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Table of Contents

1.0    EXECUTIVE SUMMARY			1
1.1    Introduction			1
1.2    Property Description and Location			1
1.3    Project geology			1
1.4    Exploration Drilling and Sampling			1
1.5    Mineral Resource Estimate			1
1.6    Mineral Processing and Metallurgical Testing			1
1.7    Mining Operations			1
1.8    Processing and Recovery Options			2
1.9    Infrastructure			3
1.10    Capital and Operating Costs			3
1.11    Economic Analysis			4
1.12    Environmental Studies, Permits and Social or Community Impact			6
1.13    Conclusions			7
1.14    Recommendations			7
2.0    INTRODUCTION AND TERMS OF REFERENCE			11
2.1    Terms of References			11
2.2    Sources of Information			11
2.3    Units and Currency			13
3.0    RELIANCE ON OTHER EXPERTS			18
3.1    Expert Reports			18
3.2    Reliance on External Reports and Exports			18
4.0    ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY			19
4.1    Location			19
4.2    Property Description and Mineral Tenure			19
4.3    Moroccan Mining Legislation			22
4.4    Acquisition History			23
5.0    ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOLOGY			24
5.1    Accessibility			24
5.2    Infrastructure			24
5.3    Climate			25
5.4    Physiography			25
5.5    Local Resources			26
6.0    HISTORY			28
6.1    Exploration and Mining History			28
6.2    Antiquity to the sixteenth Century			28
6.3    BRPM Exploration (1956-1985)			28
6.4    Underground Development and Mining (1986–1992)			28
6.5    BRPM Exploration (1993–1998)			29
6.6    Maya Gold & Silver / Aya Gold & Silver (2013–Present)			29
6.7    Historical Production			30
6.8    Previous Mineral Resource Estimates			30
7.0    GEOLOGICAL SETTING AND MINERALIZATION			33
7.1    Regional Geology			33
7.2    Property Geology			35
7.3    Structural Geology			38

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7.4    Mineralization			41
8.0    DEPOSIT TYPES			44
9.0    EXPLORATION			45
9.1    Geochemical Sampling/Grab Samples			46
9.2    Historical Underground Sampling			46
9.3    Bulk Sample			46
9.4    Hyperspectral Surveys			47
9.5    Geophysical Surveys			49
9.6    Geological and Structural Mapping			50
9.7    Trenching			51
10.0    DRILLING			52
10.1    Diamond Drilling			53
10.2    Reverse Circulation Drilling			54
10.3    Collar Surveys			54
10.4    Downhole Surveys			54
10.5    Hole Orientation			55
10.6    Core Orientation			58
10.7    SRK Comments			59
11.0    SAMPLE PREPARATION, ANALYSIS AND SECURITY			60
11.1    Introduction			60
11.2    Core Recovery			60
11.3    Geotechnical Logging			62
11.4    Geological Logging			62
11.5    Photography			64
11.6    Density and Moisture Measurements			65
11.7    Sampling			67
11.8    Core Cutting and Drying			68
11.9    Sample Preparation			69
11.10    Sample Analysis			71
11.11    Sample Security			72
11.12    SRK Comments			72
12.0    DATA VERIFICATION			73
12.1    Introduction			73
12.2    SRK Verification			73
12.3    Verification by the Company and External Consultants			78
12.4    Aya Drillcore re-Sampling			79
12.5    Quality Assurance and Quality Control (QA/QC) Analysis			83
12.6    SRK Comments			99
13.0    MINERAL PROCESSING AND METALLURGICAL TESTING			100
13.1    Introduction			100
13.2    Results Interpretation and Conclusions			127
13.3    Future Testwork Recommendations			132
14.0    MINERAL RESOURCE ESTIMATES			134
14.1    Introduction			134
14.2    Mineral Resource Estimation Procedure			134
14.3    Database Construction, Validation and Comments			135
14.4    Exploratory Data Analysis			137
14.5    Topography and Collars Adjustments			140
14.6    Geological Modelling and Domaining			141
14.7    Lithological Model			142

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14.8    Weathering Model			142
14.9    Mineralization Model			143
14.10    Data Conditioning and Post-domaining Statistical Analysis			149
14.11    Block Model Validation			170
14.12    Mineral Resource Classification			186
14.13    Mining Depletion			188
14.14    Reconciliation			190
14.15    Economic and Technical Input Parameters			191
14.16    Mineral Resource Statement			195
14.17    Sensitivity Analysis			198
14.18    Aya Peer Review			199
14.19    Comparison to Previous Estimates			199
15.0    MINERAL RESERVE ESTIMATES			202
16.0    MINING METHODS			203
16.1    General Description of the Mining Project			203
16.2    Geotechnical			203
16.3    Hydrogeology			206
16.4    Boumadine Mining Method			206
16.5    Open Pit Mining			207
16.6    Underground Mining			215
16.7    Stockpile Management			227
16.8    Production Plan			229
16.9    Electrical Underground Distribution			245
16.10    Ventilation			247
17.0    RECOVERY METHODS			257
17.1    Process Design			257
17.2    Process Plant Description			261
17.3    Plant Consumption			267
18.0    PROJECT INFRASTRUCTURE			269
18.1    Overall Site Development			269
18.2    Site Access Road			269
18.3    Process Plant Facilities			269
18.4    Mining Facilities			272
18.5    Tailings Storage Facility			274
18.6    Water Supply			275
18.7    Water Management			276
18.8    Power Supply			276
18.9    Sewage and Solid Waste Management			276
18.10    Security			277
18.11    Accommodations			277
18.12    Administration and Plant Buildings			277
18.13    Concentrate Transportation and Loadout			277
19.0    MARKET STUDIES AND CONTRACTS			278
19.1    Market Studies			278
19.2    Commodity price projections			278
19.3    Contracts			278
20.0    ENVIRONMENTAL STUDIES, PERMITS, AND SOCIAL OR COMMUNITY IMPACTS			280
20.1    Project Setting			280
20.2    Environmental and Social Studies			280

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20.3    Permitting and Applicable Standards			281
20.4    Key Environmental and Social Issues			281
21.0    CAPITAL AND OPERATING COSTS			284
21.1    Capital Cost Estimate			284
21.2    Operating Cost Estimate			294
22.0    ECONOMIC ANALYSIS			305
22.1    Introduction			305
22.2    Fiscal and Economic Parameters			305
22.3    Project Timing			306
22.4    Financial Modelling			306
22.5    Results			312
22.6    Conclusion			318
23.0    ADJACENT PROPERTIES			319
24.0    OTHER RELEVANT DATA AND INFORMATION			321
24.1    Project Execution Schedule			321
24.2    Pyrite Treatment			321
25.0    INTERPRETATION AND CONCLUSIONS			323
25.1    Geology			323
25.2    Mining			324
25.3    Processing			325
25.4    Infrastructure			326
25.5    Environmental and Social			327
25.6    Capital Cost Estimate			327
25.7    Operating Cost Estimate			328
25.8    Economic Analysis			328
25.9    Conclusion			329
26.0    RECOMMENDATIONS			330
26.1    Summary			330
26.2    Geology			330
26.3    Mining			332
26.4    Processing			334
26.5    Tailings Storage Facility			335
26.6    Environmental and Social			336
26.7    Project Next Phases			336
27.0    REFERENCES			337
28.0    CERTIFICATES			341
Appendix A    QAQC CRM PLOTS			356
Appendix B    STATISTICAL VALIDATION - UNCAPPED VS CAPPED COMPOSITES			427
Appendix C    STATISTICAL VALIDATION _ CAPPED COMPOSITES VS BLOCK MODEL			465

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List of Tables

Table 1-1    SRK Mineral Resource Statement for the Boumadine Project, Morocco, as at an effective date of 28 February 2026			1
Table 1-2    Capital Expenditures including Sustaining and Closing Costs			4
Table 1-3    Operating Costs for First 5 Years of Operation and for LOM			4
Table 1-4    Key Economic Results			5
Table 1-5    Additional Economic Model Outputs			5
Table 2-1    Qualified Persons Responsible for this Report			12
Table 2-2    Terminology and Abbreviations			13
Table 2-3    Unit Measurement Abbreviations			17
Table 4-1			21
Table 6‑1    BRPM 1998 Boumadine MRE (not compliant with modern CRIRSCO reporting codes)			30
Table 6‑2    GoldMinds Geoservices Inc 2019 Boumadine MRE (effective date 15 February 2019)			31
Table 6‑3    P&E Mining Consultants Inc. 2024 Boumadine MRE (effective date 23 February 2024)			32
Table 6‑4    Aya Gold & Silver Inc. 2025 Boumadine MRE (effective date 24 February 2025)			32
Table 9‑1    Summary of Aya Database provided by Aya to SRK			45
Table 10‑1    Table showing drilling contractor and downhole survey used between 2022 and 2026 (as provided by source Aya)			55
Table 13-1    Summary of Testwork Programs			100
Table 13-2    Head Analysis of Fresh Rock and Tailings Samples			102
Table 13-3    Roast-CIL Results			105
Table 13-4    POX-CIL Results			105
Table 13-5    CIL Results Summary			107
Table 13-6    Composite Sample Head Analysis			107
Table 13-7    Composite Sample Mineralogy			108
Table 13-8    Locked Cycle Test Conditions			111
Table 13-9    Locked Cycle Test Results Summary			112
Table 13-10    Pyrite Concentrate Analysis			113
Table 13-11    Direct CIL Extraction Results			113
Table 13-12    Pyrite Concentrate Albion Testing Results			113
Table 13-13    CIL Leaching Results for Albion Residues			114
Table 13-14    Bio-Oxidation Testing Results			114
Table 13-15    Bio-Oxidation Residue CIL Results			114
Table 13-16    Roast Testing Results			115
Table 13-17    Roast Residue CIL Results			115
Table 13-18    Locked Cycle Test Results Summary			119
Table 13-19    Roast Residue CIL Results			120
Table 13-20    SGS Bulk Flotation Results			122

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Table 13-21    Elemental Analysis Results of Pyrite Concentrate (%)			122
Table 13-22    Cyanidation Pre-Treatment Evaluation			125
Table 13-23     Comparison of Pyrite Treatment Methods			129
Table 14‑1    Summary of the Aya drillhole database within the Boumadine Mining Licence			136
Table 14‑2    Final estimation domains			148
Table 14‑3    Selected top cuts on raw assay data			150
Table 14‑4    Selected top cuts on metal accumulation data (grade x thickness) from SRK’s visual validation			152
Table 14‑5    SRK modelled variogram parameters			159
Table 14‑6    Selected domains with soft boundaries applied			162
Table 14‑7    Boumadine block model framework			164
Table 14‑8    Fixed plunge applied to the estimation group			166
Table 14‑9    Comparison of Mines Production Record and Depleted Block Model			191
Table 14‑10    NSR Zinc Concentrate Calculation			193
Table 14‑11    NSR Lead Concentrate Calculation			194
Table 14‑12    NSR Pyrite Concentrate Calculation			194
Table 14‑13    Total NSR by Element			194
Table 14‑14    Cut-off value calculation			194
Table 14‑15    SRK Mineral Resource Statement for the Boumadine Project, Morrocco, as at an effective date of 28 February 2026			196
Table 16-1    Summary of OSA recommendations.			203
Table 16-2    Recommended stope sizing constraints by stope face and rock type.			204
Table 16-3 Pillar Dilution Based by Mining Sequence.			205
Table 16‑3    NSR Cut-off Calculations General Assumptions			207
Table 16‑4    Whittle Parameters			207
Table 16-6    Open Pit Design Parameters			213
Table 16-7    Stope Design Parameters			217
Table 16-8    Deswik Stope Optimizer (DSO) Output by Zones D			217
Table 16-9    Development Headings Profiles			218
Table 16‑10    Mine Dewatering			220
Table 16‑11    Operating mobile equipment units – Annual requirements			224
Table 16‑12    Maximum Labour Requirements			226
Table 16-13    Metal Prices and Smelter Terms by Concentrate			227
Table 16-14    NSR-Mine Factor Derivation			228
Table 16-15    NSR-Pyrite Calculation Parameters			228
Table 16-16    Stockpile Classification Criteria			229
Table 16‑17    Open Pit Life-Of-Mine Production by Stockpile			232
Table 16‑18    Open Pit Life-Of-Mine Production by Zone			232
Table 16-19    Development Performance Rate			236
Table 16‑20    Life-of-Mine Underground Development Plan			237
Table 16‑21    Stope Production Cycle Parameters			240
Table 16‑22    Underground Production Plan Sorted by Stockpile Category			241
Table 16‑23    Underground Production Plan Sorted by Mining Zone			241

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Table 16‑24    Boumadine Life-of-Mine Production Plan			243
Table 16-25    Surface Substations by Portal			245
Table 16-26    Electrical and Communication Infrastructure Unit Costs			246
Table 16-27    Diesel equipment air demand			248
Table 16-28    Task resource definitions			248
Table 16-29    Peak airflow by zone, including 15% leakage			249
Table 16-30    Main fan operating points and installation configuration			253
Table 16-31    Summary of raises by dimensions and length			253
Table 16-32    Energy Consumption of Main Fans			254
Table 16-33    Energy Consumption of Auxiliary and Ramp Development Fans			256
Table 17-1    Key Process Design Criteria			260
Table 17-2    Daily Water Consumption (Nominal)			267
Table 17-3    Annual Energy Consumption			267
Table 17-4    Annual Consumption of Reagents and Consumables			268
Table 18-1    Design Criteria			274
Table 21-1    Capital Estimate Summary			284
Table 21-2    Currency Exchange Rate			284
Table 21-3    Capital Cost Estimate - Mining			285
Table 21-4    General Facilities Cost			286
Table 21-5    Open Pit Facilities Cost			286
Table 21-6    Underground Development Cost			287
Table 21-7    Mobile Equipment			287
Table 21-8    Underground Facilities Cost			287
Table 21-9    Main fans cost			288
Table 21-10    Process Plant and Infrastructure Capital Cost Breakdown			288
Table 21-11    Derivation of Quantities			290
Table 21-12    Supply Cost Source			290
Table 21-13    Owners Project Cost Total			291
Table 21-14    LoM Capital Costs (Million MAD)			292
Table 21-15    Capital Expenditures including Sustaining and Closing Costs			292
Table 21-15    Operating Cost Estimate - Mining			295
Table 21-16    Open-Pit Unit Costs			295
Table 21-17    Open-Pit Owner’s Team			296
Table 21-18    Underground Variable Costs			296
Table 21-19    Underground Fixed Costs			296
Table 21-20    Mobile Maintenance Costs			297
Table 21-21    Process Plant Operating Cost Summary (US$, Q3 2025, +50 / -30%)			301
Table 21-22    LOM Blend - Power Cost Summary (US$, Q3 2025, +50 / -30%)			301
Table 21-23    Consumables Cost Summary (US$, Q3 2025, +50 / -30%)			302
Table 22-1    Base Case Prices and Sensitivity Ranges			306
Table 22-2    Key PEA Assumptions and Economic Parameters			307
Table 22-3    Cash Flow Model			309

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Table 22-4    Key Economic Results			312
Table 22-5    Additional Economic Model Outputs			313
Table 22-6    Project Economics Sensitivity - Gold and Silver Price			315
Table 22-7    Base Case Prices and Sensitivity Ranges			316
Table 22-8    Gold and Silver Price Based IRR - Pre-Tax			316
Table 22-9    Gold and Silver Price Based IRR - Post-Tax			316
Table 22-10    Gold and Silver Price Based NPV5% Pre-Tax in $M			316
Table 22-11    Gold and Silver Price Based NPV5% Post-Tax in $M			317
Table 22-12    CAPEX and OPEX Based IRR - Pre-Tax			317
Table 22-13    CAPEX and OPEX Based IRR - Post-Tax			317
Table 22-14    CAPEX and OPEX Based NPV5% Pre-Tax in $M			318
Table 22-15    CAPEX and OPEX Based NPV5% Post-Tax in $M			318
Table 25-1    Key Economic Results			328
Table 26-1    Budget for Advancing Project to Next Stages			330

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List of Figures

Figure 1-1    Boumadine Project - Annual Free Cash Flow Profile			6
Figure 4-1    Location of Boumadine (WGS84 UTM Zone 30N)			19
Figure 4‑2    Boumadine Mining Licence in relation to Aya surrounding Licences			22
Figure 5‑1    View looking across the Boumadine project area			25
Figure 5‑2    View looking over historical Boumadine mine			26
Figure 7‑1    Regional geology of the Anti-Atlas and the Ougnat Massif.			34
Figure 7‑2    Simplified geological map of the Boumadine Project			37
Figure 7‑3    Conceptual structural model proposed by Gaboury (2022)			39
Figure 7‑4    WorldView-3 alteration map and structural interpretation presented by Gaboury, 2022.			40
Figure 7‑5    Examples of sulphide mineralization textures from the Boumadine Project.			42
Figure 7‑6    Conceptual model illustrating five configurations of sulphide mineralization at Boumadine			43
Figure 9‑1    Inclined image showing the exploration database provided by Aya plotted in relation to the Boumadine Mining Licence (black), collars coloured by Year with traces coloured grey			46
Figure 9‑2    Map of the 2022 Hyperspectral Survey Completed on Boumadine, (source Aya 2025)			48
Figure 9‑3    2022 VTEM Magnetic Survey Completed at Boumadine (source Aya 2025)			50
Figure 10‑1    DD drill rig set up at the Project during the SRK QP site visit			53
Figure 10‑2    Plan view showing the modelled wireframes (green) vs the drillhole traces at Boumadine			56
Figure 10‑3    Cross-sections A and B showing interpreted mineralized zones and drillhole traces.			57
Figure 10‑4    Core orientation procedure using the ACTx tool.			58
Figure 11‑1    Histogram showing the core recovery data provided to SRK for review			61
Figure 11‑2    Image of BOU-DD26-817, taken during the SRK site visit			61
Figure 11‑3    Aya core logging facility			63
Figure 11‑4    Aya core photography stand, with inlay image of drillcore image in Imago Capture software			65
Figure 11‑5    Aya density measurement station			66
Figure 11‑6    Active sampling of drillcore at Aya core processing facility			67
Figure 11‑7    Sampled drillcore with the sample identification and individual QR code denoted on the bag			68
Figure 11‑8    Aya core saw facility			69
Figure 11‑9    Rocklabs sample crusher			70
Figure 11‑10    Pulverized samples being prepared for shipment to Afrilab analytical laboratory in Marrakech			71
Figure 12‑1    Drillhole BOU-DD95-404 verified by SRK in the field			75
Figure 12‑2    Gossanous outcrop observed at surface during the high level geological mapping undertaken by SRK			76

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Figure 12‑3    Fire assay furnaces (A) and the inductively coupled plasma (ICP) analytical instrument (B) at the Afrilab laboratory, Marrakech, Morocco.			77
Figure 12‑4    Scatter plot showing Aya resampled drillcore (Au g/t)			80
Figure 12‑5    Scatter plot showing Aya resampled drillcore (Ag g/t)			80
Figure 12‑6    Scatter plot showing Aya resampled drillcore (Fe %)			81
Figure 12‑7    Scatter plot showing Aya resampled drillcore (Pb%)			81
Figure 12‑8    Blank performance between 2018 and 2026 (A= Au, B = Ag, C= Cu, D= Pb and E = Zn)			85
Figure 12‑9    Ag multi CRM plots (A = full range, B = rescaled Y axis)			88
Figure 12‑10    Au multi CRM plots (A = full range, B = rescaled Y axis)			89
Figure 12‑11    Cu multi CRM plots (A = full range, B = rescaled Y axis)			90
Figure 12‑12    Pb multi CRM plots (A = full range, B = rescaled Y axis)			91
Figure 12‑13    Zn multi CRM plots (A = full range with sample reporting 1299% Zn omitted, B = rescaled Y axis)			92
Figure 12‑14    Field duplicate performance between 2018 and 2026 (A = Au and B = Ag)			94
Figure 12‑15    Coarse rejects (DUP and DUP_REJECT) inserted by Aya into the sample stream for Au and Ag.			96
Figure 12‑16    Umpire duplicate samples sent to ALS by Aya for Au and Ag			98
Figure 13-1    Metallurgical Sampling Locations			101
Figure 13-2    Locked Cycle Test Flowsheet			110
Figure 13-3    Cleaner Flotation Flowsheet for Separating Arsenopyrite / Pyrite			117
Figure 13-4    Correlation Between Analyzed and Calculated Sulphur Content			121
Figure 13-5    Au and Ag Leach Rates at Different Roasting Temperatures			123
Figure 14‑1    Plan view of drillhole traces (coloured grey) and collars used in the 2026 MRE(			137
Figure 14‑2    Box plot showing raw Ag, Au and Zn grades with Aya lithological logging codes			139
Figure 14‑3    Boumadine topographic wireframe in plan view coloured by elevation and drillhole collars			141
Figure 14‑4    Cross section (30m slice) through weathering model,			143
Figure 14‑5    Plan view showing location of the North, Central and Suther Wireframes			145
Figure 14‑6    Boumadine modelled wireframes in (light brown) in relation to the drillhole traces (grey)			146
Figure 14‑7    Comparison of MS and SMS structural measurements			147
Figure 14‑8    Histogram of raw Ag g.t sample data for group XX in south			151
Figure 14‑9    Histogram of raw Zn % sample data for group XX in central area			152
Figure 14‑10    Histogram of raw Ag metal accumulation data for domain 1091 in South			153
Figure 14‑11    Histogram of raw Au metal accumulation data for domain 2091 in Central			154
Figure 14‑12    Example capped metal accumulation histogram for Au in domain 2091			155
Figure 14‑13    Example capped metal accumulation histogram for Ag in domain 2091			156
Figure 14‑14    Directional variograms for Pb (grade x thickness) for group C1 (central)			157
Figure 14‑15    Directional variograms for true thickness (TT) for group N2_Tizzi (north)			158
Figure 14‑16    Directional variograms for Zn (grade x thickness) for group S2 (south)			159
Figure 14‑17    Example of QKNA results for block sizes			163
Figure 14‑18    Vertical section (looking north) of a modelled domains (red and green) with local search ellipses shown (blue) to demonstrate application of LVA			166

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Figure 14‑19    Coded density database showing the 13 SGDRY samples excluded from the analysis			168
Figure 14‑20    Plot showing regression applied to density database used to determine density when no SGDRY values have been recorded			169
Figure 14‑21    Long section of domain 1031			172
Figure 14‑22    Long section of domain 1091			173
Figure 14‑23    Long section of domain 2091			174
Figure 14‑24    Long section of domain 3103			175
Figure 14‑25    Long section of domain 3193			176
Figure 14‑26    Long section of domain 3193			177
Figure 14‑27    Long section of domain 2091			178
Figure 14‑28    Cross section in the Southern area through 317165 X and 3475490 Y (azimuth 340° dip 90°)			179
Figure 14‑29    Cross section in the Central area through 316911 X and 3477037 Y			180
Figure 14‑30    North south cross section in the Northern area through 315420. X and 3478310 Y			181
Figure 14‑31    Mean metal accumulation (Au) composite Grades (blue line) compared to Block Estimate Grades			183
Figure 14‑32    Mean metal accumulation (Au) composite Grades (blue line) compared to Block Estimate Grades			183
Figure 14‑33    Mean metal accumulation (Ag) composite Grades (blue line) compared to Block Estimate Grades			184
Figure 14‑34    Mean metal accumulation (Zn) composite grades (blue line) compared to block estimate grades			185
Figure 14‑35    Density composite grades (blue line) compared to block density values			186
Figure 14‑36    View of the block model coloured by Mineral Resource classification			188
Figure 14‑37    Underground mining as-builts (stopes = blue and development = black) provided by Aya (location central area)			189
Figure 14‑38    Cross-section looking north at mined as-builts provided by Aya (grey)			190
Figure 14‑39    Boumadine Grade Tonnage Curves for Open Pit Indicated and Inferred Material			198
Figure 13‑40    Boumadine Grade Tonnage Curves for Underground Indicated and Inferred Material			198
Figure 16-1    Whittle Re-Blocked Model Parameters			208
Figure 16-2    Whittle Initial Slope Set Settings			209
Figure 16-3    North Pit Whittle Shells Iteration			210
Figure 16-4    North Pit Whittle Shell 14			210
Figure 16-5    Central Pit Whittle Shells Iteration			211
Figure 16-6    Central Pit Whittle Shell 16			211
Figure 16-7    South Pit Whittle Shells Iteration			212
Figure 16-8    South Pit Whittle Shell 11			212
Figure 16-9    Haul Road Schematic (a) and (b)			213
Figure 16-10    Boumadine Pit Locations			214
Figure 16-11    Boumadine Pit Center Phases – Longitudinal Section			215
Figure 16‑12    Boumadine Underground Mine Design – Longitudinal View			215
Figure 16‑13    Modified Avoca Mining Method Illustration – Longitudinal Section			216
Figure 16‑14    Modified Avoca Mining Method – Stope Cycle			216
Figure 16‑15     Typical Level Layout			219

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Figure 16‑16     Dewatering Network Schematic Central Zone			220
Figure 16‑17    Cement Slurry Process and Transportation Underground			221
Figure 16‑18     Mixing Sump Schematic			222
Figure 16‑19    Stope Backfill Placement in Poor Ground Conditions			223
Figure 16‑20    Stockpile Classification Parameters			229
Figure 16‑21    Open Pit Sequence: Surface view every 2 year: (a) 2030 (b) 2032 (c) 2034 (d) 2036 (e) 2038 (f) 2039			231
Figure 16-22    Mineralized Material Mined per Year and per Pits			233
Figure 16-23    Waste Mined per Year per Pits			233
Figure 16-24    Open Pit Life-of-Mine Production Plan and NSR			234
Figure 16-25    Development Heading (South Zone Portal 1)			236
Figure 16‑26    Primary UG Mining Area in the South Zone			238
Figure 16‑27    Secondary UG Mining Area in the Center Zone			239
Figure 16‑28    Remaining UG Mining Area in the North Zone			239
Figure 16-29    Underground Mineralized Material Mined by Zones			242
Figure 16-30    Underground Mineralized Material Mined and NSR			242
Figure 16‑31    Boumadine Life-Of-Mine Production Plan Breakdown by Mining Methods			244
Figure 16‑32    Boumadine LOM Production Plan Tonnes and NSR			244
Figure 16-33    Life of mine airflow by zone Portal 1 to Portal 5			249
Figure 16-34    Zones and locations of exhaust fans			250
Figure 16-35    Longitudinal view of South 1 zone showing three exhaust raises and three intakes			250
Figure 16-36    Longitudinal view of South 2 zone showing exhaust raise and intake portal			251
Figure 16-37    Longitudinal view of South 3 zone showing three exhaust raises and three intakes			251
Figure 16-38    Longitudinal view of Center (Portal 4) zone showing two exhaust raises and two intakes			252
Figure 16-39    Longitudinal view of North (Portal 5) zone showing four exhaust raises and four intakes			252
Figure 16-40    Auxiliary Ventilation on Ramp Development Zone Centre			256
Figure 16-41    Auxiliary Ventilation on Production Levels Zone Centre			256
Figure 17-1    Simplified Process Flowsheet			258
Figure 17-2    Simplified Process Flowsheet (Continued)			259
Figure 18-1    Overall Site Plan			270
Figure 18-2    Process Plant Layout			271
Figure 18-3    Raw Water Sourcing			275
Figure 21-1    Capital Cost Summary – Mining			285
Figure 21-2    Operating Cost Summary – Mining			295
Figure 21-3    Underground Labour Profile			297
Figure 21-4    Processing Plant Operating Cost Breakdown			300
Figure 22-1    Boumadine Project - Annual Free Cash Flow Profile			312
Figure 22-2    Annual Free Cash Flow			314
Figure 23-1    Location of the Imiter Mine			320
Figure 24-1    Project Execution Schedule			321
Figure 25-1    Boumadine Project - Annual Free Cash Flow Profile			329

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1.0EXECUTIVE SUMMARY

1.1Introduction

This Technical Report (the “Report”) presents an Updated Preliminary Economic Assessment ("PEA") for the Boumadine Project, Kingdom of Morocco. Boumadine is a polymetallic deposit, the commodities of interest are gold (Au), silver (Ag), Zinc (Zn), Lead (Pb). The Report has an effective date of September 9, 2026.

To support the preparation of the Updated PEA, Aya retained a number of independent engineering and consulting firms to provide specialized technical input. The roles and responsibilities of the consultants contributing to this report are summarized below:

•SRK Consulting (UK) Ltd prepared the updated Mineral Resource Estimate (MRE) for the project.

•CCE Mining developed the mining methods, mining capital cost and operating cost estimates.

•Lycopodium (Americas) Ltd. ("Lycopodium") developed PEA-level designs for the process plant and general on-site infrastructure, developed the consolidated cost estimates and reviewed and signed off on the financial model that was prepared by the Company.

•RockEng Inc. prepared the mining geotechnical evaluations

•The power supply line was estimated by the Office National de l'Électricité et de l'Eau Potable ("ONEE") and reviewed by Lycoodium

•SLR Consulting conducted environmental studies and permitting reviews.

•Epoch Resources (Pty) Ltd prepared and reviewed the Tailings Storage Facility evaluation

The following QP's contributed in the preparation of this report:

•Preetham Nayak, P.Eng., Lycopodium

•Zunedbhai Shaikh, P.Eng., Lycopodium

•Ruan Venter, P.Eng., Lycopodium

•Paul Gauthier, P.Eng., CCE Mining

•Alex Pheiffer, PrSciNat, SLR Consulting France SAS

•George Papageorgiou PrEng, PhD, MSc, BSc, Eng (Civil), Wits, Epoch Resources (Pty) Ltd

•Guy Dishaw, P.Geo., SRK Consulting (UK) Ltd

SRK has been requested by Aya Gold and Silver Inc. (Aya, hereinafter also referred to as the Company or the Client) to prepare an update to the Mineral Resource Estimate (MRE) for the Boumadine polymetallic deposit (Boumadine or the Project). This MRE represents the first Mineral Resource Estimate that SRK has been commissioned to undertake for the Boumadine Project. The MRE presented in this report has been prepared using the terminology, guidelines and definitions provided in the Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Definition Standards for Mineral Resources and Mineral Reserves (May 2014). This technical report is prepared in accordance with National Instrument 43-101 – Standards of Disclosure for Mineral Projects (NI-43-101). The Mineral Resource Statement presented in this report has been prepared and approved by Mr Guy Dishaw, P.Geo., a Qualified Person (QP) in accordance with the CIM Definition Standards. Mr Dishaw is a member of the Association of Professional Engineers and Geoscientists of Saskatchewan. Guy Dishaw (QP) completed a site visit to the Project between 23 and 28 April 2026 to review the local geology, mineralization, drilling, sampling practices, core handling, core orientation procedures, the Company's data

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management systems and visited the African Laboratory for Mining and Environment (Afrilab) sample preparation onsite and the assay laboratory located in Marrakech. The MRE presented in this report incorporates the latest phase of exploration and infill drilling. The drilling database was provided to SRK on the 16 April 2026 and was last updated with assays issued from the laboratory on the 28 February 2026. Accordingly, the Mineral Resource Statement (MRS) has an effective date of 28 February 2026 (hereinafter referred to as the “2026 MRE”) which in turn reflect various technical and economic conditions at the time of writing.

Aya engaged Lycopodium to prepare an updated PEA for the Boumadine Project. The PEA has been completed in accordance with the Canadian disclosure requirements of National Instrument 43-101 (NI 43-101) and the associated requirements of Form 43-101F1.

1.2Property Description and Location

Boumadine is located within the Drâa-Tafilalet Region of southeastern Morocco within the Central Anti-Atlas Mountain Belt. It is located approximately 340 km southeast of Rabat, 220 km east northeast of Ouarzazate, 290 km east of Marrakech and approximately 80 km southwest of Errachidia.

The Project is located within the Boumadine Mining Licence (LE-383661), which is held by Boumadine Global Mining SA (BGM). BGM is owned 85% by Aya Gold & Silver Inc. (Aya) and 15% by the Office National des Hydrocarbures et des Mines (ONHYM). The Boumadine Mining Licence covers an area of approximately 31.7 km², which expired in May 2026. Aya have informed SRK that they are currently in the processes of renewing the Licence (no change to the extent or conditions) for an additional 10 years. SRK understand that there is no reason why the renewal will not be successful. The Company also holds an additional 13 Mining Licences and 19 Exploration Permits within the surrounding region.

1.3Project geology

The Boumadine Project is hosted within Neoproterozoic volcanic and volcano-sedimentary rocks of the Tiddiline Formation (TTF) and comprises a structurally controlled polymetallic hydrothermal system developed within a shallow volcanic environment. Mineralization is principally controlled by steeply dipping northwest- to north-northwest-trending fault corridors that are spatially associated with felsic subvolcanic intrusions and hydrothermal breccias. Hydrothermal alteration is extensive and comprises quartz-sericite-pyrite (phyllic) alteration within the felsic volcanic sequence and propylitic alteration within the underlying mafic volcanic rocks, characteristic of a long-lived hydrothermal system that evolved during repeated episodes of fault movement and magmatism.

Mineralization comprises steeply dipping massive to semi-massive sulphide veins, hydrothermal breccias, replacement sulphides and pyrite stockwork developed within broad alteration halos. The highest-grade mineralization is typically associated with massive sulphide veins dominated by pyrite, sphalerite, galena, arsenopyrite and chalcopyrite, with surrounding zones of disseminated sulphides and quartz-carbonate veining. Recent drilling and geological studies have demonstrated that mineralization extends into both the felsic and mafic volcanic successions and has led to the refinement of the structural interpretation of the deposit, providing increased confidence in the geological model and continuity of the principal mineralized corridors. Although the precise genetic classification of the deposit remains under investigation, the MRE is based on the observed geological, structural and mineralization characteristics and does not depend on any specific genetic model.

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1.4Exploration Drilling and Sampling

A total of 940 drillholes, comprising 356,323.1 m of drilling, are located within the Boumadine Mining Licence. Of these, 855 drillholes, totalling 319,464.8 m, were completed to support infill of key mineralized areas and extensions to the known mineralization, while the remaining 85 drillholes relate to prospecting within the Mining Licence and did not specifically target the known Boumadine mineralization.

Drillhole spacing is variable across the Project. However, the majority of the modern drilling used in the 2026 MRE was initially completed on nominal 100-200 m-spaced inclined drill fences, with subsequent infill drilling undertaken on approximately 50 m spacing. The infill drilling program is ongoing at the time of writing. In certain locations, drillhole spacing is closer owing to variations in drillhole inclination or where additional drilling was planned to provide increased drill coverage.

In the opinion of SRK, the sampling procedures used by the Company generally conform to industry best practices and the resulting drilling pattern is sufficiently dense to interpret the geometry, geological boundaries and sulphide mineralization with an appropriate level of confidence.

1.5Mineral Resource Estimate

In summary, for the 2026 Mineral Resource update, SRK has completed the following:

•compilation, review and validation of the geological, drilling, sampling and analytical databases;

•review of geological mapping, satellite/hyperspectral imagery, geophysical dataset, structural mapping and oriented core data to characterise the geological controls on the orientation, continuity and geometry of the mineralization;

•construction of weathering and oxidation (redox) model;

•construction of geological and estimation domain wireframes (mineralization wireframes), guided by geological mapping, structural measurements from oriented drill core, historical mine plans and as-built depletion surveys, together with Aya’s previous geological interpretations;

•conditioning of the assay database, compositing, metal accumulation (grade × thickness) calculations and assessment of high-grade outliers;

•geostatistical analysis and variogram modelling using the metal accumulation data;

•construction of a three-dimensional block model using parent block dimensions of 5 m × 15 m × 15 m within the mineralized zones and 90 m × 90 m × 90 m for the surrounding waste;

•interpolation of Ag, Au, Cu, Pb, Zn and true thickness (TT) using Ordinary Kriging (OK), whilst Fe and As were estimated using Inverse Distance Weighting (IDW2). All metals were estimated using the metal accumulation method;

•estimation of density using single drillhole composites representing the across-strike mineralized intersections, with factoring applied within the transitional weathering domain. Density values were based on measured density data supplemented by density values derived from regression analysis undertaken by SRK;

•visual and statistical validation of the block model against the input composite data, including comparison of estimated metal accumulation values, density composites and back-calculated grades against the corresponding input data

•classification of the block model into the Indicated and Inferred Mineral Resource categories based on data quality, geological continuity and drillhole spacing;

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•assessment of the reasonable prospects for eventual economic extraction (RPEEE) through the determination of appropriate reporting cut-off values and the generation of conceptualised open pit shells and underground reporting shapes;

•reporting of the Mineral Resources in accordance with the definitions, terminology and guidelines of the CIM Definition Standards for Mineral Resources and Mineral Reserves (May 2014);

•comparison of the current Mineral Resource Statement (MRS) with the previous MRS prepared by Aya.

SRK considers that the Boumadine mineralization has reasonable prospects for economic extraction with respect to both open-pit and underground mining methods.

The Boumadine MRS has an effective date of 28 February 2026, and has been depleted to reflect the current understanding of the historical mining. The Mineral Resources are limited to classified material contained within conceptual open pit shell reported above the NSR reporting cut-off of 60 USD/t and, below the pit shell, to classified material contained within grade shells (following the removal of isolated shapes) based on a 110 USD/t NSR cut-off, that satisfy a 1.5 m minimum mineable thickness. The Mineral Resources are reported within the Boumadine Mining Licence on an in-situ, undiluted basis are shown in Table ES 1 below.

The Company’s Boumadine Mining Licence permit expired in May 2026. Aya have informed SRK that they are currently in the processes of renewing the Licence (no change to the extent or conditions) for an additional 10 years. Whilst the licence renewal approval is currently pending, the Company has informed SRK that they see no reason why this will not be successful.

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Table 1-1    SRK Mineral Resource Statement for the Boumadine Project, Morocco, as at an effective date of 28 February 2026

			Cut-off (NSR USD/t)			Tonnes (Kt)			Average Grade																					Contained Metal
									Ag (g/t)			Au (g/t)			Cu (%)			Pb (%)			Zn (%)			AgEq (g/t)			AuEq (g/t)			Ag (koz)			Au (koz)			Cu (kt)			Pb (kt)			Zn (kt)			AgEq (koz)			AuEq (koz)
Pit-Constrained
Indicated			60			6,639			115.5			2.2			0.10			0.96			2.21			390.8			4.2			24,653			463			6			64			147			83,423			896
Inferred			60			19,497			57.8			2.0			0.07			0.62			1.54			298.7			3.2			36,218			1,280			14			120			301			187,258			2,011
Underground
Indicated			110			1,943			143.0			1.1			0.06			1.19			2.11			319.6			3.4			8,931			69			1			23			41			19,963			214
Inferred			110			25,868			56.0			1.6			0.08			0.56			1.17			251.9			2.7			46,605			1,372			21			145			303			209,475			2,249
Total
Indicated			60/110			8,582			121.7			1.9			0.09			1.01			2.19			374.7			4.0			33,585			532			8			87			188			103,386			1,110
Inferred			60/110			45,365			56.8			1.8			0.08			0.58			1.33			272.0			2.9			82,823			2,653			35			265			604			396,733			4,260

In reporting the Mineral Resource Statements, SRK notes the following:

•The Mineral Resources have an effective date of 28 February 2026. Mr Guy Dishaw, P.Geo., is the Qualified Person (QP) responsible for the Mineral Resource Estimate in accordance with the Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Definition Standards for Mineral Resources and Mineral Reserves (2014) and CIM Estimation of Mineral Resources and Mineral Reserves Best Practice Guidelines (2019). Mr Dishaw is a member of the Association of Professional Engineers and Geoscientists of Saskatchewan.

•The Mineral Resource Estimate was prepared by a team of consultants from SRK under the supervision of the QP.

•Mineral Resources are reported in situ and undiluted and are constrained within the Boumadine Mining Licence.

•Mineral Resources have been depleted to reflect SRK's current understanding of historical underground mining completed up to mine closure in 1992 as well as surficial artisanal workings.

•Commodity prices of US$2,800/oz Au, US$30/oz Ag, US$4.60/lb Cu, US$1.00/lb Pb and US$1.20/lb Zn were used to establish reasonable prospects for eventual economic extraction.

•The NSR was calculated as: NSR (US$/t) = (Pb (%) × 12.87) + (Zn (%) × 14.12) + (Au (g/t) × 70.27) + (Ag (g/t) × 0.75) + (Cu (%) × 66.95) − 6.75.

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•Silver equivalent (AgEq) and gold equivalent (AuEq) grades are based solely on metal prices and metallurgical recoveries and do not incorporate the full range of factors included in the NSR calculation, including treatment charges, payabilities, penalties, royalties and other smelter charges. Consequently, equivalent grades were not used for optimization or application of reporting cut-offs and are provided for reference only.

•AgEq = Ag (g/t) + (Au (g/t) × Au price/gram × Au recovery) / (Ag price/gram × Ag recovery) + Zn (%) × Zn price/lb × Zn recovery / (Ag price/gram × Ag recovery) × 685.7147973 + Pb (%) × Pb price/lb × Pb recovery / (Ag price/gram × Ag recovery) × 685.7147973 + Cu (%) × Cu price/lb × Cu recovery / (Ag price/gram × Ag recovery) × 685.7147973.

•AuEq = Au (g/t) + (Ag (g/t) × Ag price/gram × Ag recovery) / (Au price/gram × Au recovery) + Zn (%) × Zn price/lb × Zn recovery / (Au price/gram × Au recovery) × 685.7147973 + Pb (%) × Pb price/lb × Pb recovery / (Au price/gram × Au recovery) × 685.7147973 + Cu (%) × Cu price/lb × Cu recovery / (Au price/gram × Au recovery) × 685.7147973.

•Open-pit Mineral Resources are reported within an optimized pit shell using an NSR cut-off value of US$60/t. The optimization and reporting assumptions include ore mining costs of US$2.0/t, waste mining costs of US$1.5/t Drilling and blasting costs (grade control) US$1.1/t, US$2.0/t surface haulage and rehandling costs, processing and shipping costs of US$49/t, G&A costs of US$6/t and an overall pit slope angle of 47°.

•Underground Mineral Resources are reported using a rounded NSR cut-off value of US$110/t, based on total underground mining costs of US$55.6/t, processing and shipping costs of US$48.9/t and G&A costs of US$6.7/t. The underground Mineral Resources demonstrate sufficient continuity suitable for long-hole underground mining methods. Reporting shapes apply a minimum true mining width of 1.5 m, with isolated or discontinuous volumes considered unlikely to support potential underground extraction excluded from the reported Mineral Resource.

•Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability.

•The QP is unaware of any known environmental, permitting, legal, title, taxation, socio-economic, marketing, political or other relevant factors that could materially affect the Mineral Resource Estimate.

•Tonnages are reported in metric units and grades are reported in grams per tonne (g/t) for Au and Ag and percent (%) for Cu, Pb and Zn. Tonnages, grades and contained metal quantities have been rounded appropriately. Rounding may result in apparent differences between totals and the sums of individual values; such differences are not considered material.

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1.6Mineral Processing and Metallurgical Testing

Based on the 2022 and 2024 metallurgical testwork completed by SGS Canada Inc., the process plant design for the Project considers a conventional flotation flowsheet. Lead, zinc and pyrite concentrates will be recovered and sold to market.

OMC used its testwork database to infer missing data and to recommend further testwork for subsequent studies. While both BWi values from testwork fell outside the database range for other similar resources in the area, the average BWi fell at the 81st percentile of the range, aligning well with OMC’s typical strategy of designing to the 85th percentile hardness. As such, the average of both test results was used in the comminution circuit design (13.1 kWh/t). The remainder of the comminution circuit design criteria were estimated from OMC’s database.

The conditions tested during the 2022 and 2024 SGS locked cycle testing campaigns were used as the basis for the concentrate recovery circuit design, since the testing incorporated recycles at steady state and produced concentrates of acceptable grade and recovery. Using locked cycle data provides a higher design certainty since it more closely represents plant operation by the incorporation of recycles and since it validates the process conditions at steady state.

The design criteria derived from the 2022 and 2024 locked cycle testing includes:

•General circuit design (i.e. orientation, number of stages, implementation of regrinding).

•Reagent addition.

•Concentrate mass pulls, grades and recoveries.

•Residence times.

Next steps include continued refinement of the metallurgical testwork, particularly comminution, variability and dewatering.

1.7Mining Operations

The Boumadine deposit is divided into three main zones: North, Central and South. The mine plan combines open pit (OP) and underground (UG) mining. OP mining starts in 2029, and UG development starts two years later, in 2031, with first UG mineralized material in 2032. This phased approach generates early cash flow from the OP mine while deferring the capital-intensive UG development.

Over the life of mine (LOM), 41.2 Mt of mineralized material will be mined at an average NSR value of US$193/t, of which OP mining contributes 22.4 Mt (54.2%) and UG mining 18.9 Mt (45.8%). The LOM production plan extends over fifteen years, including pre-production, with OP mining from 2029 to 2039 and UG mining through 2043. Production during the final four years comes exclusively from the UG mine. Mineralized material is classified into three stockpiles (Low Grade, LG Pyrite and HG Pyrite) based on NSR values, and the stockpiles are used to regulate the tonnage and grade of plant feed at the nominal throughput of 8,000 t/d.

1.7.1OP Mining

Pre-production activities, including pre-stripping and the construction of roads, pads and waste rock storage areas, will start in 2029 during the project construction phase. Three pits are planned. The Central Pit, which contains the highest-grade OP mineralized material, is the main pit and will be mined in four phases over approximately eleven years. The South Pit will be mined in two phases to supplement Central Pit production, and the North Pit, made up of three sub-pits, will be mined from 2034 to 2036. Total material movement peaks at 55 Mt/y, and the LOM stripping ratio is approximately 23:1.

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Pit shells were optimized in GEOVIA Whittle and designed in Deswik using the geotechnical recommendations of RockEng. The first two benches are 10 m high to account for near-surface weathering, followed by 20 m (double 10 m) benches in rock, with geotechnical slope breaks every 100 vertical metres. OP mining will be carried out by a contractor using a conventional truck and shovel fleet, based on 200 t class hydraulic shovels (CAT 6020 or similar) paired with 90 t haul trucks (CAT 777 or similar). Haul ramps are designed at a maximum grade of 10%.

1.7.2UG Mining

The UG mine extends to approximately 640 m below surface over approximately 5.6 km of strike. The mineralized zones are steeply dipping and average 1 to 4 m in width. The UG mine will be accessed by ramps from five surface portals: Portals 1 to 3 in the South Zone, Portal 4 in the Central Zone and Portal 5 in the North Zone. Development starts at Portal 1 in 2031, followed by Portal 4 in 2034, Portals 3 and 2 in 2036 and Portal 5 in 2038. Initial UG production focuses on the higher-grade South Zone, and UG production reaches approximately 2.0 Mt/y from 2038.

Based on the geometry of the deposit and the geotechnical assessment, the selected mining method is modified Avoca, a longitudinal longhole retreat method mined bottom-up within each panel and backfilled with rock fill (RF) and cemented rock fill (CRF). Stopes are designed with a 20 m level spacing, a 20 m strike length and a minimum mining width of 2.0 m, with 0.5 m of hanging wall dilution and 95% mining recovery. Mineralized material and waste will be hauled to surface by 33 t trucks (Epiroc MT436B or similar) via the ramps. Ventilation raises to surface will be excavated by raise boring, and internal raises by drop raising or raise boring.

Capital lateral and vertical development will be performed by a contractor. Sill development and all stoping activities will be performed by owner crews.

1.8Processing and Recovery Options

The process plant design for the Project is based on a conventional flotation flowsheet. Lead, zinc and pyrite concentrates will be recovered and sold to market.

The plant design criteria were selected to minimize operating costs and maximize the use of proven technology. The key criteria for equipment selection are suitability for duty, safety, reliability, and ease of maintenance.

The process plant design comprises the following unit operations:

•Single stage primary crushing with a jaw crusher to produce an 80% Passing (P80) crushed product size of 136 mm.

•Crushed rock stockpile with a live capacity of approximately 8,115 tonnes to provide 24 hours of live storage. During extended periods of primary crusher equipment maintenance, additional crushed material inventory can be generated in the weeks leading up to the planned shutdown by dozing crushed material from this stockpile to the area adjacent to the stockpile. This material can then be reclaimed during the shutdown by front-end loader to feed the grinding circuit.

•SAB type grinding circuit consisting of a SAG mill and a ball mill to produce a ground product with a P80 of 58 microns (µm), with hydrocyclones for particle size classification.

•Lead rougher flotation, classification, regrinding and cleaner flotation to produce a lead concentrate of 29.6% grade. The lead concentrate will be thickened, filtered, and bagged prior to shipment.

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•Zinc rougher flotation, classification, regrinding and cleaner flotation to produce a zinc concentrate of 57.4% grade. The zinc concentrate will be thickened and filtered for stockpiling prior to shipment.

•Pyrite rougher flotation to produce a pyrite concentrate with 4.2 Au g/t and 81.0 Ag g/t. The pyrite concentrate will be thickened and filtered for stockpiling prior to shipment.

•Tailings thickening and storage in a tailings management facility.

Run-of-mine (ROM) material will be delivered to a feed bin by mine truck to supply the crushing plant. The crushed product will be stockpiled before entering the grinding circuit, which will include a SAG mill, ball mill, and hydrocyclones for particle size classification. The ground product will feed the lead rougher flotation circuit. The lead rougher concentrate will proceed to a classification and regrind circuit, followed by three cleaner flotation stages and a cleaner-scavenger stage. The final lead concentrate will be thickened, filtered and bagged for shipment.

The lead rougher tails and cleaner-scavenger tails will feed the zinc rougher flotation circuit. The zinc rougher concentrate will follow a similar regrind and cleaning process, including three cleaning stages and a cleaner-scavenger stage. The final zinc concentrate will be thickened, filtered, and stockpiled for shipment.

The zinc rougher tails and cleaner-scavenger tails will feed the pyrite rougher flotation circuit. The pyrite rougher tails will be thickened and pumped to the flotation tailings management facility. The pyrite concentrate will be thickened, filtered and stockpiled for shipment. Filtrate and thickener overflows from all circuits will return to the process water pond.

1.9Infrastructure

Infrastructure on the Project includes an integrated mining and processing operation including open pits, an underground mine, and a central processing plant supported by crushing, flotation, stockpiling, and concentrate handling facilities. Site access is via the National 10 highway with upgraded roads connecting all major areas.

Key infrastructure includes the ROM pad, workshops, laboratories, security and administration buildings, and extensive haul and service roads. Waste rock storage, a lined TSF with staged raises, and on-site water harvesting and treatment support operational needs. Power is provided through a new 225 kV ONEE supply to a 40 MVA substation. Additional services include fuel storage, sewage treatment, waste management, security, limited accommodation, and concentrate transport to Nador-West Port.

1.10Capital and Operating Costs

1.10.1Capital Costs

The capital cost estimates for the Boumadine Project PEA was compiled by Lycopodium. All costs are expressed in United States Dollars (US$) unless otherwise stated. The capital cost estimate was developed in Q3 2025 and escalated to Q3 2026 pricing and deemed to have an overall accuracy of +50%/-30%. The capital cost estimate conforms to AACE International (Association for the Advancement of Cost Engineering) Class 5 estimate standards as prescribed in recommended practice 47R11.

The total initial capital cost for the Boumadine Project is $463 M and the LOM sustaining cost is $507 M. The initial capital and sustaining capital, for the Project is summarized in Table 1-2 by major area.

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Table 1-2    Capital Expenditures including Sustaining and Closing Costs

Capital Expenditures ($M)			Initial			Sustaining			Total
Direct Costs			299			507			806
Open Pit Mining			52			59			112
Underground Mining			0			407			407
Processing Plant			175			0			175
Shipping Infrastructure			11			0			11
Electrical Line			18			0			18
Raw Water Supply			33			0			33
Tailings Storage Facility			9			30			40
TSF Closure Costs			0			11			11
Indirects Costs			65			0			65
Subtotal			364			507			871
Contingency			99			0			99
Total			463			507			970

1.10.2Operating Costs

The operating cost estimate conforms to a preliminary economic assessment level estimate with a +50/-30% accuracy. The operating cost estimate was developed in Q4 2025 using data from projects, studies, and previous operations from Lycopodium, CCE, and Aya.

The LOM average unit operating cost is $67.61/t and the total LOM cost including cash cost and sustaining cost is $122.55/t. Table 1-3 provides a summary of the operating costs for the Project.

Table 1-3    Operating Costs for First 5 Years of Operation and for LOM

Item			Units			Year 1-5			LOM
Mining			$/t milled			51.47			43.70
Processing			$/t milled			18.08			17.96
G&A			$/t milled			5.58			5.45
Tailings, Environmental and Water Management			$/t milled			0.47			0.50
Total On-Site Operating Cost			$/t milled			75.60			67.61
Transportation			$/t milled			38.63			34.33
Royalties			$/t milled			10.65			8.00
Mining Tax			$/t milled			0.34			0.32
Total Cash Cost			$/t milled			125.22			110.26
OP Sustaining Capital			$/t milled			3.08			1.44
UG Sustaining Capital			$/t milled			9.41			9.85
TSF Sustaining Capital			$/t milled			0.95			1.00
Total Costs Including Sustaining			$/t milled			138.67			122.55

1.11Economic Analysis

The economic analysis was performed assuming a 5% discount rate typical for silver-gold projects. Cash flows have been discounted to the start of construction, assuming that the project execution decision will be made, and major project financing will be carried out at this time.

The pre-tax NPV discounted at 5% is $4.1B, with a pre-tax IRR of 97% and payback period of 0.7 years. On an after-tax basis, the NPV discounted at 5% is $3.5B, with an after-tax IRR of 93%, and payback period of 0.7

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years. Cumulative after-tax unlevered free cash flow totals $4.7B. Tax calculations are based on Aya’s understanding of current Moroccan tax regulations as of the effective date of this report.

Readers are cautioned that the Updated PEA is preliminary in nature. It includes Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves and there is no certainty that the PEA will be realized. Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability.

A summary of the project economics is listed in Table 1-4 and Table 1-5, and after-tax free cash flow is shown graphically in Figure 1-1.

Table 1-4    Key Economic Results

Project Economics			Units			Base Case
						Pre-tax			Post-tax
NPV5%			$B			4.1			3.5
IRR			%			97		%	93		%
Payback			Years			0.7			0.7
NPV: Capex1			-			8.6			7.6
Revenue LOM			$B			11.0			-
Avg. Annual Revenue			$M/y			765			-
EBITDA LOM			$B			6.1			-
Avg. Annual EBITDA			$M/y			427			-
Cumulative FCF LOM			$B			5.5			4.7
Avg. Annual FCF			$M/y			381			327

1.NPV:Capex ratio is the ratio of Net Present Values, discounted at 5%, to the initial capital expenditure.

Table 1-5    Additional Economic Model Outputs

Item			Units			Year 1-5			LOM
Processed Grade
Au			g/t			2.05			1.77
Ag			g/t			101.6			63.5
Zn			%			2.04			1.37
Pb			%			0.87			0.58
Au-eq			g/t			4.11			3.10
Ag-eq			g/t			283			216
Recoveries
Au rec (%)			%			96.1%			96.1%
Ag Rec (%)			%			96.4%			96.4%
Zn rec (%)			%			74.7%			74.7%
Pb Rec (%)			%			82.0%			82.0%
Metal Production
Gold Production			koz			877			2,254
Silver Production			koz			43,683			81,167
Zinc Production			Mlbs			467			931
Lead Production			Mlbs			219			431
Gold-Equivalent Production (Au-eq)			koz			1,740			3,889
Silver-Equivalent Production (Ag-eq)			koz			121,283			272,219
Operating Cost per Ounce

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Item			Units			Year 1-5			LOM
Total Cash Costs1			$/oz AuEq			998			1,169
Total AISC2			$/oz AuEq			1,105			1,300

1.Cash costs include mine-site operating costs such as mining, processing, and direct site G&A, as well as product shipping, royalties and mining taxes.

2.AISC is calculated as the sum of treatment and refining charges, onsite operating costs, sustaining capital costs, and closure costs, divided by the quantity of ounces equivalent produced.

Figure 1-1    Boumadine Project - Annual Free Cash Flow Profile

1.12Environmental Studies, Permits and Social or Community Impact

The biophysical and social characteristics of the planned mining area are defined by its semi-arid climate and relatively rural and remote nature. Urban settlements are concentrated along main transport routes with villages located along valleys with access to water resources. Climatic conditions influence local ecosystems and land uses including agriculture, concentrated in areas of water availability.

Limited environmental studies have been completed in the past. As part of the environmental and social work currently being undertaken, an environmental baseline sampling program was initiated by BGM in September 2024 and a high-level screening for critical biodiversity habitat was undertaken by SLR in September 2024.

The baseline sampling, covering soils, air quality and water, was undertaken by AfriLabs, a Moroccan-based laboratory specializing in mineral and environmental analysis.

Additional baseline surveys have been completed during the course of 2026 covering social, biodiversity, soil, water, air quality and noise. The results are being processed to inform the international ESIA process outlined below. Studies covering the topics of biodiversity, hydrology, hydrogeology, soils and sediments, air quality, noise, GHG, climate change, stakeholder engagement, labour and working conditions, community health and safety, and other social elements are ongoing.

The Project will require national permits to support the development of the mine and the requirement to carry out an Environmental Impact Assessment (EIA) under Moroccan Law. The EIA under Moroccan Law, prepared by NOVEC, was completed and submitted in July 2026 to the “Centre Régional d'Investissement” (CRI) (Regional Investment Center). In addition to meeting national permitting requirements, the Project is being undertaken to meet international ESIA lender standards.

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A conceptual mine closure plan governed by Aya’s Environmental Policy will be prepared as part of the ESIA.

1.13Conclusions

The Project is at an advanced stage of exploration and geological understanding, supported by drilling and sampling, geological mapping, orientated core, geophysical surveys, historical mine plans as well as previous geological studies. In addition, no significant issues were identified with the analysis of the quality control and assurance (QAQC) programs. These studies provide confidence to the local scale geometry of the mineralization and grade distributions in the resultant Mineral Resources and has allowed SRK to produce a robust geological model with over 158 individual domains modelled. However, there are areas of lower geological confidence, classified as Inferred Mineral Resources and there are some domains SRK was not able to classify, which may be subject to further revision in the future. SRK notes that the there is a degree of uncertainty associated with depletion survey as-builts and that these underground workings are currently inaccessible.

SRK has declared a Mineral Resource Statement for the Boumadine Project. The Mineral Resources are limited to material contained within the conceptualised open pit shell above the NSR reporting cut-off of 60 USD/t and, below the pit shell, to material contained within grade shells (following the removal of isolated shapes) based on a 110 USD/t NSR cut-off that satisfy a minimum 1.5 m minable thickness. The Mineral Resources are reported within the Boumadine Mining licence on an in-situ, undiluted basis and amount to:

•Indicated Mineral Resources (open pit and underground) of some 8,582 kt, at a mean grade of 121.7 g/t Ag, 1.9 g/t Au, 0.09 % Cu, 1.01 % Pb and 2.19 % Zn; and

•Inferred Mineral Resources (open pit and underground) of some 45,365 kt, at a mean grade of 56.8 g/t Ag, 1.8 g/t Au, 0.08 % Cu, 0.58 % Pb and 1.33 % Zn.

The Updated PEA provides a base case assessment for developing the Boumadine mineral resource by conventional open pit and underground mining methods, and a crushing-milling-flotation processing plant with a tailings dam.

The Updated PEA economic analysis shows the Boumadine Polymetallic Project has an after-tax NPV5% of $3.5B, IRR of 93%, and a payback period of 0.7 years. The PEA economics support a decision to continue to advance the Project and carry out additional detailed studies, including pre-feasibility level work and completion of a feasibility study.

1.14Recommendations

Additional expenditures are recommended by the Boumadine Report Authors for the following activities:

•Advance metallurgical testwork, including spatial composites and variability samples.

•Undertake detailed studies for infrastructure, including TSF design, raw water sourcing, electrical line connection and concentrate logistics studies.

•Update and upgrade mining and processing designs through pre-feasibility and feasibility level designs.

•Conduct hydrogeology and geotechnical studies to improve design parameters for mining.

•Complete the ESIA, which will inform a comprehensive Environmental and Social Management and Monitoring Plan to be developed for the Project.

SRK considers there to be opportunities to further improve confidence in the geological interpretation, MRE, and supporting datasets for the Boumadine Project. SRK recommends that Aya consider the following:

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•In relation to drilling and geological understanding:

•All drillholes should be designed to intersect the interpreted mineralized wireframes as close to perpendicular as practicable, thereby maximising representativity and location accuracy.

•Continue the ongoing infill drilling program to improve geological and grade continuity that support the conversion of Inferred Mineral Resources to the Indicated category.

•Complete additional twinned diamond drillholes (DD) at representative locations across the deposit, targeting selected existing RC, MP and DD holes completed between 2018 and 2026. Twin holes should be selected to provide representative coverage of the principal mineralised domains, mineralisation styles and grade ranges, one of the target areas should be where different drilling and sampling methods overlap. The twinning program should be used to assess drilling and sampling precision, identify any systematic bias between drilling methods, quantify short-scale geological and grade variability (including the nugget effect), and provide additional information to support the confidence assigned to the current MRE.

•Complete additional step-out diamond drilling to test the strike and depth extensions of the known massive sulphide domains, where this remains open.

•Aya should investigate whether whole rock geochemistry will help with the identification and standardisation of the lithologies encountered in the drilling.

•Include total sulphur (S%), and where practicable sulphide/sulphate sulphur analyzes, within future drilling programs to improve/confirm definition of the oxidation profile and weathering boundary.

•In relation to sampling and core management:

•Record the sample weight for every RC drillhole interval (including rejects) and compare this with the theoretical sample mass based on the logged lithology to identify intervals with poor recovery that may require re-drilling.

•Photograph all available drill core and RC chip trays and incorporate these records into the Imago database to establish a comprehensive digital photographic archive.

•Include all drillcore photographs taken prior to 2024 into the Imago database.

•Continue to maintain and periodically review sampling and QA/QC procedures as exploration progresses.

•In relation to density:

•Expand the density measurement program, particularly within the mineralized wireframes and transitional (weathered) mineralized zones, to improve the representativeness of the density dataset and reduce reliance on regression-derived density estimates.

•Collect density measurements from drillholes that currently contain no density information and consider obtaining additional measurements from archived drill core where appropriate.

•Consider the use of plastic film or paraffin wax coatings during density determinations to minimize moisture ingress for friable or fractured core.

•Undertake a representative moisture content sampling program using fresh drill core collected from across the Project to improve future mining and processing studies, particularly for the Feasibility Study.

•In relation to QA/QC:

•Expand the duplicate QA/QC database to include the complete analytical suite and include the following assays obtained as a minimum, Cu, Pb and Zn prior to the next Mineral Resource update.

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•Introduce pulp duplicate samples into the routine QA/QC program.

•Consider introducing a commercially certified blank material or undertake a round-robin program to better characterise the locally sourced blank material.

•Review duplicate records where assay values are missing or recorded as zero and update the database against the original laboratory certificates where appropriate.

•Investigate and correct anomalous CRM assay values, including the Zn CRM records noting as being in excess of 100%.

•Consider discontinuing OREAS 100A, OREAS 354 and OREAS 994 as these are not certified CRMS for base metal analysis using Aqua regia or for Au and Ag using fire assay analysis.

•Investigate the small number of CRM sample swaps/mislabels that remain within the QA/QC database, resulting either from the mislabelling of original drill core samples or from other QA/QC samples (including CRMS) being incorrectly assigned, and resolve these where possible.

•Continue the routine umpire laboratory program and consider increasing the proportion of samples submitted to the umpire laboratory, including representative samples from the 2024 drilling program.

•Where practicable, ensure that future umpire analyzes use analytical methods with reporting limits comparable to those of the primary laboratory.

•In relation to surveying:

•Incorporate the planned high-resolution pre-mining drone survey into future geological models and associated mining studies.

•Where safe and practicable, resurvey historical underground workings (once mining begins) and surface artisanal excavations to improve confidence in future depletion models.

•In relation to Resource Estimation:

•Develop a 3D lithological model, subject to the availability of sufficient geological data, to improve geological understanding and support future Mineral Resource, Mineral Reserve, geotechnical, and mine planning studies.

•Undertake a change-of-support study following additional infill drilling to optimize the parent block dimensions and improve future Mineral Resource estimates.

•Incorporate a Mineable Stope Optimizer (MSO) into future Mineral Resource estimates using parameters representative of the proposed underground mining method for Resource reporting.

•Aya should expand its metallurgical test work program to include the weathered/transitional material in order to demonstrate this has reasonable prospects for eventual economic extraction.

•In relation to Net Smelter return Calculation and Equivalent cut-off value calculations:

•Aya should review the NSR royalty calculation in future updates so that royalties are applied directly to the relevant concentrate streams and remain linked to the underlying value drivers.

•SRK recommends that Aya model concentrate penalties and deductions separately from treatment charges, using concentrate grades as inputs.

•Aya should consider incorporating freight costs within the NSR calculation in future updates, rather than operating costs, to link these costs directly to the saleable concentrate mass.

•Review the concentrate mass calculation methodology to account for the sequential processing flowsheet. Specifically, Pb concentrate should be based on plant feed grades, Zn concentrate

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should be based on Pb-circuit tailings grades, and pyrite concentrate should be based on Zn-circuit tailings grades. This would affect the calculated mass pull and therefore shipment costs. The difference is expected to be immaterial for Pb and Zn concentrates but may be more relevant for pyrite concentrate.

•SRK recommends calculating the equivalent Ag and Au grades using NSR factors, opposed to the current approach which uses metal prices and metallurgical recoveries only. This approach will avoid any inconsistency with the NSR framework already in place.

SRK has not prepared a detailed exploration budget for the Project but has reviewed the budget and associated schedules provided by Aya. These principally relate to infill drilling intended to upgrade portions of the currently classified Inferred Mineral Resources to Indicated Mineral Resources and support the proposed 15- year life-of-mine plan. Based on Aya's proposed program, approximately 326 000 m of additional drilling is anticipated, with estimated drilling and sampling costs of approximately US$ 65 million. This cost estimate excludes the multidisciplinary Feasibility Study (FS) planned by Aya for 2027.

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2.0INTRODUCTION AND TERMS OF REFERENCE

2.1Terms of References

This Technical Report on the Boumadine (also known as Boumaadine, Boumâadine, and Bou Madine) Deposit was prepared by Aya Gold and Silver Inc. (“Aya” or the “Company”) at the request of Mr. Benoit La Salle, President & CEO of Aya Gold & Silver Inc. Aya is a public, TSX listed company trading under the symbol “AYA”, with its head office located at: Suite 132, 1320 Graham, Ville Mont-Royal, Québec, H3P 3C8. This Technical Report has an effective date of September 9, 2026.

This Technical Report (“the Report”) has been prepared to provide a fully compliant NI 43-101 Technical Report and Preliminary Economic Assessment ("PEA") of the existing mineralization at the Boumadine Deposit (or the “Boumadine Property” or the “Boumadine Mining License”), located in the Kingdom of Morocco. The Boumadine Mining License is a joint venture owned by Aya (85%) and l’Office National des Hydrocarbures et des Mines (“ONHYM”) (15%) of the Kingdom of Morocco. Aya (known as Maya Gold & Silver at the time of the Acquisition Agreement) and Office National des Hydrocarbures et des Mines (“ONHYM”) signed an agreement on October 9, 2012, for the acquisition of the Boumadine Mining License. A new Moroccan company, Boumadine Global Mining (“BGM”), was created with Aya (85%) and ONHYM (15%) as shareholders. The mining title of the Boumadine Mining License was transferred to BGM by ONHYM.

The name change from Maya Gold & Silver Inc. to Aya Gold & Silver Inc. was announced in a press release dated July 30, 2020. The Updated Mineral Resource Estimate reported herein is based on up-to-date drilling results and appropriate metal pricing, and is fully conformable to the “CIM Standards on Mineral Resources and Reserves – Definitions (2014) and Best Practices Guidelines (2019)”, as referred to in National Instrument (“NI”) 43-101 and Form 43-101F, Standards of Disclosure for Mineral Projects.

2.2Sources of Information

2.2.1Site Visit

In accordance with NI43-101, Guy Dishaw (QP) of SRK completed a site visit to the Project between 23 and 28 April 2026, accompanied by Pierre Pihouée of Aya. The site visit took place during the course of the commission to gain an understanding of the local geology, styles of mineralization, drilling practices, core handling procedures, core orientation methods, sampling protocols, and the data management systems employed by the Company, as well as to interview Project personnel. During the site visit, Guy Dishaw also undertook high-level mapping of mineralization exposed at surface, verified the collar positions of 14 drillholes, and inspected key Project infrastructure, including areas of historical artisanal mining. In addition, Guy visited the Company's sample preparation laboratory on site and the assay laboratory operated by Afrilab, located in Marrakech.

Observations made during the site visit were consistent with the geological interpretations adopted for the Mineral Resource Estimate (MRE) and provided confidence that the drilling, sampling, and data collection procedures employed by the Company are suitable for supporting an MRE.

Two additional independent Qualifed Persons have visited the site. Cortney Palleske, M.A.Sc., P.Eng., visited the Boumadine Property from June 5 to 7, 2024 and July 7 to 16, 2026, to meet with project geological team, inspect core, and tour the property as part of the mining geotechnical evaluation program. Alex Pheiffer, PrSciNat, visited the site from September 1 to 3, 2024 and January 19 to 20, 2026 as part of the Environmental and Social Impact Assessment ("ESIA"), met with the geology team and toured site to assess proposed locations of infrastructure.

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2.2.2Additional Information Sources

This Report is based, in part, on internal Company technical reports, database, and maps, published government reports, company letters and memoranda, and public information as listed in the references (Section 27) of this Report. Several sections from reports authored by other consultants have been directly quoted or summarized in this Technical Report, and are cited where appropriate.

The Authors of this Report have used selected portions or excerpts from material contained in the following NI 43-101 compliant Technical Reports, which are publicly available on SEDAR+ under Aya’s profile:

•Boily, M. 2014. The Boumadine Polymetallic (Au, Ag, Zn, Pb, Cu) Deposit, Errachidia Province, Kingdom of Morocco. Prepared for Maya Gold & Silver Inc. by Boily, Michel. Effective Date April 2, 2014. Issue Date April 2, 2014. 217 pages.

•GoldMinds Geoservices Inc. 2019. NI 43-101 Technical Report Preliminary Economic Assessment [of the] Boumadine Polymetallic Deposit, Kingdom of Morocco. Prepared for Maya Gold & Silver Inc. by Duplessis, C., Rachidi, M., Dufort, D. and Rousseau, G. Effective Date: April 24, 2019. Issue Date: May 24, 2019. 304 p.; and

•P&E Mining Consultants Inc. 2024. NI 43-101 Technical Report and updated Mineral Resource Estimate of the Boumadine Polymetallic Project, Kingdom of Morocco. Prepared for Aya Gold & Silver Inc. by Stone, W., Brown, F., Barry, J., Yassa, A., Feasby, G. and Puritch, E. Effective Date: May 08, 2024. Issue Date: May 31, 2024. 214 p.;

•Aya Gold & Silver 2025. NI 43-101 Technical Report and Updated Mineral Resource Estimate of the Boumadine Polymetallic Project, Kingdom of Morocco. Prepared for Aya Gold & Silver Inc. by Lalonde, D, and Pérez, P. Effective Date: February 24, 2025. Issue Date : March 31, 2025. 193 p.

•Aya Gold & Silver 2025. NI 43-101 Technical Report and Preliminary Economic Assessment for the Boumadine Polymetallic Project, Kingdom of Morocco. Prepared for Aya Gold & Silver Inc. by Nayak, P., Venter, R., Shaikh, Z., Berson, B., Pheiffer, A., Papageorgiou, G., Puritch, E., Yassa, A., Brown, F., Barry, J., Stone, W., and Palleske, C. Effective Date: November 4, 2025. Issue Date : December 18, 2025. 326 p.

Since 2024, Aya has continued extensive mineral exploration and drilling programs. The results of these programs were disclosed publicly in numerous Aya press releases and are summarized in Sections 9 and 10 of this Report.

The Authors and Co-authors of each section of this Report are listed in Table 2-1, who in acting as Qualified Persons as defined by NI 43-101, take responsibility for those sections of this Report as outlined in the “Certificate of Author” included in Section 28 of this Report.

Table 2-1    Qualified Persons Responsible for this Report

Qualified Person			Company			Sections of Technical Report
Preetham Nayak, P.Eng.			Lycopodium (Americas) Ltd.			2, 3.2, 19, 21.1.1, 21.1.3, 21.1.6-21.1.10, 21.2.4, 22, 24.1 and Co-Author 1, 25, 26, 27
Ruan Venter, P.Eng.			Lycopodium (Americas) Ltd.			13, 17, 21.2.2 - 21.2.3, 24.2 and Co-Author 1, 25, 26, 27
Zuned Shaikh P.Eng.			Lycopodium (Americas) Ltd.			18.1 to 18.3, 18.6 to 18.13, and Co-Author 1, 25, 27
Paul Gauthier, P.Eng.			CCE Mining			16.1, 16.3-16.10, 18.4, 21.1.2, 21.2.1 and Co-Author 1, 25, 26, 27
Alex Pheiffer, PrSciNat			SLR Consulting France SAS			20 and Co-Author 1, 25, 26, 27
George Papageorgiou PrEng, PhD, MSc, BSc, Eng (Civil), Wits			Epoch Resources (Pty) Ltd			18.5, 21.1.4-21.1.5, 21.2.5, and Co-Author 1, 25, 26, 27

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Qualified Person			Company			Sections of Technical Report
Guy Dishaw, P.Geo.			SRK Consulting (UK) Ltd			3.1, 4-12, 14, 23.1 and Co-author 1, 14, 25, 26, 27
Cortney Palleske, M.A.Sc., P.Eng.			RockEng Inc.			16.2 and Co-Author 25, 27

2.3Units and Currency

In this Technical Report, all currency amounts are stated in US dollars (“$”) unless otherwise stated. Commodity prices are typically expressed in US dollars (“US$”) and will be duly noted where appropriate. Quantities are generally stated in Système International d’Unités (“SI”) metric units including metric tons (“tonnes”, “t”) and kilograms (“kg”) for weight, kilometres (“km”) or metres (“m”) for distance, hectares (“ha”) for area, grams (“g”) and grams per tonne (“g/t”) for metal grades. Platinum group metal (“PGM”), gold and silver grades may also be reported in parts per million (“ppm”) or parts per billion (“ppb”). Copper metal values are reported in percentage (“%”) and parts per billion (“ppb”). Quantities of PGM, gold and silver may also be reported in troy ounces (“oz”), and quantities of copper in pounds (“lb”). Abbreviations and terminology are summarized in Table 2-2 and units in Table 2-3.

Grid coordinates for maps are given in the UTM WGS 84 Zone 30 R or as longitude and latitude.

Table 2-2    Terminology and Abbreviations

Abbreviation			Meaning
$			dollar(s)
°			degree(s)
°C			degrees Celsius
<			less than
>			greater than
%			percent
3-D			three-dimensional
AAS			atomic absorption spectrometry
Actlabs			Activation Laboratories Ltd.
Ag			silver
AgEq			silver equivalency
ALS			ALS Laboratory (Australian Laboratory Services), ALS Limited
ANCOLD			Australian National Committee on Large Dams
As			arsenic
Au			gold
AuEq			gold equivalency
avg			average
Aya			Aya Gold & Silver Inc.
Ba			barium
BGM			Boumadine Global Mining
BRGM			Bureau des Recherches Géologiques et Minières
BRPM			Bureau de Recherches et de Participations Minières
BWI			ball mill work index
CaO			calcium oxide
Cd			cadmium
CDA			Canadian Dam Association
CGG			Compagnie Générale de Géophysique

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Abbreviation			Meaning
CIL			carbon-in-leach
CIM			Canadian Institute of Mining, Metallurgy and Petroleum
cm			centimetre(s)
CN			cyanide
Co			cobalt
Company, the			Aya Gold & Silver Inc.
CRI			Comité Régional d’Investissement (centre of investment)
CRM(s)			certified reference material(s)
CSA			Canadian Securities Administrators
CSAMT			controlled-source audio magnetotelluric survey
Cu			copper
d			day
D80			80% passing size
DAR			Direction des Affaires Rurales
Deposit, the			Boumadine Deposit
E			east
EBRD			European Bank for Reconstruction and Development
ESMS			environmental and social management system
ESIA			Environmental and Social Impact Assessment
FA			fire assay
Fe			iron
g			gram
G&A			General and administration
g/L			grams per litre
g/t			grams of metal per tonne
GISTM			Global Industry Standard on Tailings Management
GoldMinds			GoldMinds Geoservices Inc.
ha			hectare(s)
HC			Hot Curing
HSEC			Health and Safety, Environment and Community
ICOLD			International Commission on Large Dams
ICP			inductively coupled plasma
ICP-OES			inductively coupled plasma-optical emission spectroscopy
ID			identification
ID2			inverse distance squared
IFC			International Finance Corporation
ISO			International Organization for Standardization
ISO/IEC			International Organization for Standardization/International Electrotechnical Commission
k			thousand(s)
kg			kilogram(s)
kg/t			kilogram(s) per tonne
km			kilometre(s)
km2			square kilometre(s)
koz			thousands of ounces
kt			kilotonne(s), thousands of tonnes
kWh/t			kilowatt hour per tonne
L			litre(s)

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Abbreviation			Meaning
lb			pound(s) (weight)
LB			Lime Boiling
LCT(s)			locked cycle testwork(s)
level			mine working level referring to the nominal elevation (m RL), e.g. 4285 level (mine workings at 4285 m RL)
LoM			life of mine
M			million(s)
m			metre(s)
m3			metres cubed
m asl			metres above sea level
Ma			millions of years
MAD			Moroccan Dirham
mag			magnetic(s)
max.			maximum
Maya			Maya Gold and Silver Inc.
mbs or MBS			metres below surface
Mg			magnesium
min.			minimum
mL			millilitre(s)
mm			millimetre(s)
Mn			manganese
Moz			million ounce(s)
MRE			Mineral Resource Estimate
Mt			million(s) tonnes per year
MTEDD			le ministère de transition énergétique et durable or minister of durable energetic transition
N			total number of observations/population size
N			north
N10			National Highway 10
NaCN			sodium cyanide
Ni			nickel
NI or NI 43-101			National Instrument or National Instrument 43-101
NN			Nearest Neighbour
NPV			net present value
NSR			net smelter return
OK			Ordinary Kriging
ONEE			Office National de l'Électricité et de l'Eau Potable
ONHYM			Office National des Hydrocarbures et des Mines
ORE			ORE Research & Exploration Pty Ltd.
oz			ounce(s)
P80			80% passing size
P&E			P&E Mining Consultants Inc.
PAX			potassium amyl xanthate
PEA			Preliminary Economic Assessment
Pb			lead
P.Eng.			Professional Engineer
P.Geo.			Professional Geoscientist
POX			pressure oxidation
ppm			parts per million

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Abbreviation			Meaning
Property, the			Boumadine Property
Q1, Q2, Q3, Q4			first quarter, second quarter, third quarter, fourth quarter of the year
QA			quality assurance
QA/QC			quality assurance / quality control
QC			quality control
R2			the coefficient of determination
Report, the			this Technical Report
RQD			rock quality designation
S			south
S			sulphur
SEDAR			System for Electronic Document Analysis and Retrieval
SGS			SGS Canada Inc. / SGS Lakefield Research
Si			silicon
SIPX			sodium isopropyl xanthate
SODECAT			Société de Développement du Cuivre de l’Anti Atlas
SODIM			Société de Développement de l'industrie maricole
Sr			strontium
t			metric tonne(s)
t/m3			tonnes per cubic metre
Technical Report			(this) NI 43-101 Technical Report
TEM			Transmission Electron Microscopy
TIMA-X			Tescan Integrated Mineral Analyses
TSF			tailings storage facility
TSX-V			Toronto Venture Stock Exchange
TTF			Tamerzaga-Timrachine Formation
URSTM			Unité de recherche et de service en technologie minérale
US$			United States dollars
UTM			Universal Transverse Mercator
VTEM			Versatile Time Domain Electromagnetic (survey)
W			west
w/w			weight by weight
wireframe			polygons joined together to represent a specific 3-D domain or unit
wt%			weight percent
XRD			x-ray diffraction
Zn			zinc

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Table 2-3    Unit Measurement Abbreviations

Abbreviation			Meaning			Abbreviation			Meaning
μm			microns, micrometre			m3/h			cubic metre per hour
$			dollar			m3/s			cubic metre per second
$/t			dollar per metric tonne			m3/y			cubic metre per year
%			percent sign			mØ			metre diameter
% w/w			percent solid by weight			m/h			metre per hour
¢/kWh			cent per kilowatt hour			m/s			metre per second
°			degree			Mt			million tonnes
°C			degree Celsius			Mtpy			million tonnes per year
cm			centimetre			min			minute
d			day			min/h			minute per hour
ft			feet			mL			millilitre
GWh			Gigawatt hours			mm			millimetre
g/t			grams per tonne			Mt			million tonnes or megatonnes
h			hour			MV			medium voltage
ha			hectare			MVA			mega volt-ampere
hp			horsepower			MW			megawatts
Hz			hertz			oz			ounce (troy)
k			kilo, thousands			Pa			Pascal
kg			kilogram			pH			Measure of acidity
kg/t			kilogram per metric tonne			ppb			part per billion
kHz			kilohertz			ppm			part per million
km			kilometre			s			second
kPa			kilopascal			t or tonne			metric tonne
kt			thousands of tonnes or kilotonnes			tpd			metric tonne per day
kV			kilovolt			t/h			metric tonne per hour
kW			kilowatt			t/h/m			metric tonne per hour per metre
kWh			kilowatt-hour			t/h/m2			metric tonne per hour per square metre
kWh/t			kilowatt-hour per metric tonne			t/m			metric tonne per month
L			litre			t/m2			metric tonne per square metre
L/s			litres per second			t/m3			metric tonne per cubic metre
lb			pound(s)			T			short ton
M			million			tpy			metric tonnes per year
m			metre			V			volt
m2			square metre			W			Watt
m3			cubic metre			wt%			weight percent
m3/d			cubic metre per day			yr			year

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3.0RELIANCE ON OTHER EXPERTS

3.1Expert Reports

SRK's opinion is based on information provided by the Company through its employees. Most notably, Pierre Pihouée and David Lalonde, Aya Geology Specialists, provided insight into the geological interpretation and mineralization characteristics of the deposit based on their experience at the Project. Patrick Perez, Aya Director of Technical Services, also provided information regarding the most up-to-date technical and economic parameters used to support the reporting of the MRS.

SRK has relied upon this information and, where possible, has independently verified the data provided. In addition, SRK’s QP (Guy Dishaw) completed a site visit to review the physical evidence associated with the deposit.

SRK has not performed an independent verification of land title and tenure as summarised in Section 3 of this report. SRK did not verify the legality of any underlying agreement(s) that may exist concerning the permits or other agreement(s) between third parties but has relied on the Company and its legal advisor for land title issues. SRK was informed by the Company that there are no known litigations potentially affecting the Boumadine mining Licence, which contains the Project.

SRK is not an insider, associate, or affiliate of the Company, and neither SRK nor any affiliate has acted as advisor to the Company or its affiliates in connection with the Project. The results of the MRE, as declared by SRK are not dependent on any prior agreements concerning the conclusions to be reached, nor are there any undisclosed understandings concerning any future business dealings.

SRK would like to acknowledge the support and collaboration provided by Aya personnel for this assignment. Their collaboration was greatly appreciated and instrumental to the success of this project.

3.2Reliance on External Reports and Exports

3.2.1Logistics

The QP for Section 18.13 and Section 22 has relied upon the report titled « Étude logistique du transport en vrac de minerai », prepared by SCX Technology in October 2025. The QP has not independently verified the information contained in this study and considers such reliance reasonable for the purposes of this Technical Report.

3.2.2Marketing

Lycopodium has relied on external information for Request for Interest documents provided by Aya and reviewed by Lycopodium.

3.2.3Taxes

The Qualified Person (QP) responsible for Section 22, has relied upon information, interpretations, and assumptions related to taxation provided by Aya based on current Morocco tax regulations. Lycopodium has not independently verified the taxation inputs and considers such reliance reasonable for the purposes of the Technical Report.

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4.0ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

4.1Location

Boumadine is located within the Drâa-Tafilalet Region of southeastern Morocco within the Central Anti-Atlas Mountain Belt, a well-established metallogenic province that hosts numerous precious and base metal deposits and occurrences. It is located approximately 340 km southeast of Rabat, 220 km east northeast of Ouarzazate, 290 km east of Marrakech and approximately 80 km southwest of Errachidia (Figure 4‑1). Access to the Project is provided via the national highway network, followed by a series of regional paved and gravel roads extending to the exploration camp and deposit area.

Project coordinates are referenced using the WGS84 UTM Zone 30N coordinate system, unless otherwise stated.

The approximate centre of the deposit is located at approximately 317290E and 3476780N:

Figure 4-1    Location of Boumadine (WGS84 UTM Zone 30N)

4.2Property Description and Mineral Tenure

The Project is located within the Boumadine Mining Licence (LE-383661), which is held by Boumadine Global Mining SA (BGM). BGM is owned 85% by Aya Gold & Silver Inc. (Aya) and 15% by the Office National des Hydrocarbures et des Mines (ONHYM). The Boumadine Mining Licence covers an area of approximately 31.7 km², which expired in May 2026. Aya have informed SRK that they are currently in the processes of renewing the Licence (no change to the extent or conditions) for an additional 10 years. The Company also holds an additional 11 Mining Licences and 19 Exploration Permits within the surrounding region. Figure 4‑2 shows the

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Boumadine Mining Licence in relation to Aya permits which are located next to or adjacent to the Boumadine Mining Licence.

All Mineral Resources reported in the MRE are located within the Boumadine Mining Licence. The Boumadine Mining Licence, along with other permits held by Aya in the surrounding area, is listed in Table 4-1.

The project is divided into three zones know as North (contains Tizi and Imarirren Zones), Central and South, shown in Figure 7‑2 and Figure 14‑5 below.

Historical artisanal mining activities have been recognized at several places on the Boumadine Property. Historical mining activity at Boumadine has focused on barite and Pb veins, though the sulphide mineralization at surface has also been historically mined. The BRPM also undertook underground mining of the sulphide mineralization in the three zones (North, Central and South).

SRK has relied upon information supplied by the Company regarding the ownership, status and extent of the Mining Licence and any associated surface rights comprising the Boumadine Project. The Company has stated that the mineral tenure is in good standing. Furthermore, SRK understands that there are no known legal disputes relating to the ownership of the Mining Licence or associated rights.

SRK has not undertaken an independent legal review of the mineral tenure, licence agreements, land ownership, or surface access rights, and expresses no legal opinion regarding the validity of such rights. The information presented in this report relating to mineral tenure and property ownership has been prepared based on documentation supplied by the Company and, where appropriate, publicly available information.

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Table 4-1    Aya Mining and Exploration Permits in the Boumadine Property Area

Permit ID			Permit Type			Interest of Aya*			Area (km2)			Granted			Expires
LE-383631**			Licence d'exploitation			100% Ownership			10.34			42929			46580
LE-383657**			Licence d'exploitation			Option to own 100%			8			45948			49599
LE-383661*			Licence d'exploitation			85% ownership			31.67			42507			46158
LE-383685**			Licence d'exploitation			Option to own 100%			14.67			42904			46555
LE-383692**			Licence d'exploitation			100% Ownership			4.02			42504			46155
LE-383722**			Licence d'exploitation			Option to own 100%			9.53			45562			49213
LE-383724**			Licence d'exploitation			Option to own 100%			6.16			43064			46715
LE-383763			Licence d'exploitation			Option to own 100%			14.24			45135			48787
LE-383852**			Licence d'exploitation			Option to own 100%			9.49			43389			47041
LE-383853**			Licence d'exploitation			Option to own 100%			17.59			43455			47108
LE-383856**			Licence d'exploitation			Option to own 100%			25.87			42571			46222
LE-383874**			Licence d'exploitation			Option to own 100%			8.1			42340			45992
PE-3052**			Permis d'exploitation			Option to own 100%			11.68			41837			46219
PR-3842950**			Permis de recherche			Option to own 100%			15.57			45104			46199
PR-3843051**			Permis de recherche			Option to own 100%			16			45087			46182
PR-3843056**			Permis de recherche			Option to own 100%			15.89			45087			46182
PR-3843057**			Permis de recherche			100% Ownership			9.43			45087			46182
PR-3843146**			Permis de recherche			100% Ownership			8.54			45091			46186
PR-3843156**			Permis de recherche			100% Ownership			11.72			45091			46186
PR-3843332**			Permis de recherche			100% Ownership			4.7			45211			46306
PR-3843342**			Permis de recherche			100% Ownership			11.4			45091			46186
PR-3843370**			Permis de recherche			100% Ownership			3.99			45359			46453
PR-3843371**			Permis de recherche			100% Ownership			3.03			45359			46453
PR-3843372**			Permis de recherche			100% Ownership			7.99			45359			46453
PR-3843387**			Permis de recherche			100% Ownership			8.28			45472			46566
PR-3843388**			Permis de recherche			100% Ownership			5.73			45472			46566
PR-3843389**			Permis de recherche			100% Ownership			1.56			45472			46566
PR-3843390**			Permis de recherche			100% Ownership			14.34			45472			46566
PR-3843391**			Permis de recherche			100% Ownership			13			45472			46566
PR-3843448**			Permis de recherche			100% Ownership			12.55			45797			46892
PR-3843449**			Permis de recherche			100% Ownership			14.37			45797			46892
PR-3843450**			Permis de recherche			100% Ownership			5.59			45797			46892
PR-3843581			Permis de recherche			100% Ownership			0.86			46128			47223
PR-3843582			Permis de recherche			100% Ownership			0.43			46128			47223
PR-3843583			Permis de recherche			100% Ownership			0.39			46128			47223

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Figure 4‑2    Boumadine Mining Licence in relation to Aya surrounding Licences

4.3Moroccan Mining Legislation

Mining activities within the Kingdom of Morocco are governed principally by Law No. 33-13 relating to Mines, together with its implementing regulations, including Decree No. 2-15-807, and other applicable environmental, land access, health and safety, and administrative laws.

The Moroccan mining framework provides for exploration authorisations, research permits and mining licences, together with associated renewal and transfer procedures, environmental assessment and approval requirements, mine development and operating requirements, and closure and rehabilitation obligations, where applicable.

Under the Moroccan mining code, mining titles are distinct from surface land ownership. A research permit provides the holder with the exclusive right to undertake exploration/search activities within the defined permit area for the relevant mineral substances, subject to compliance with applicable legislation, permit conditions, and work commitments.

A mining licence may be granted following the discovery and delineation of a mineable deposit, subject to satisfaction of the applicable technical, legal, administrative, and environmental requirements as noted in the Moroccan Mining Legislation.

SRK understands, based on information provided by the Company, that the Company’s exploration and mining activities are undertaken in accordance with applicable Moroccan mining legislation. SRK has not independently provided a legal opinion on Moroccan mineral tenure or regulatory compliance.

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4.4Acquisition History

On 9 October 2012, Maya Gold & Silver Inc. (Maya) and the Office National des Hydrocarbures et des Mines (ONHYM) entered into a joint venture agreement for the acquisition, development, and exploitation of the Boumadine deposit. Under the terms of the agreement, Maya acquired an 85% interest in Mining Licence LE-383661 for total cash consideration of MAD 28 million (approximately USD 2.8 million at the time of the transaction). A new Moroccan company, Boumadine Global Mining SA (BGM), was subsequently established, with Maya and ONHYM holding 85% and 15% ownership interests, respectively.

On 31 July 2020, Maya changed its name to Aya.

The Boumadine Mining Licence was subsequently transferred by ONHYM to BGM. Under the terms of the joint venture agreement, ONHYM is entitled to a 3% royalty, while Aya is entitled to a 2.75% management fee based on BGM's sales revenue from the first year of commercial production.

Aya, through its Moroccan subsidiaries, has progressively consolidated its interest in the Boumadine Region by undertaking extensive exploration, and by acquiring surrounding Exploration Permits and Mining Licences.

Based upon documentation supplied by the Company, SRK understands that the Boumadine Project is subject to certain royalty interests and contractual agreements associated with historical ownership of the Property. The Company has represented that these agreements are well understood and have been appropriately considered within previous economic evaluations of the Project. SRK has not independently reviewed the contractual agreements relating to royalties or other commercial encumbrances and therefore expresses no opinion regarding their legal interpretation or enforceability.

From the information available, SRK is not aware of any royalty or encumbrance that would materially affect the reporting of the Mineral Resources presented herein.

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5.0ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOLOGY

5.1Accessibility

The Boumadine Project is located within Errachidia Province in the Draa-Tafilalet Region of southeastern Morocco, on the northwestern flank of the Ougnat Massif in the Anti-Atlas Mountains. The Project is readily accessible from the regional road network, being located via National Highway N10. The nearest town is Tinejdad, located approximately 16 km north of the Project, whilst the village of Bouyoud is situated approximately 4 km east of the Project. Access from Tinejdad to the Project is provided by a combination of paved and well-maintained gravel roads suitable for year-round access.

5.2Infrastructure

The Boumadine Project benefits from existing infrastructure developed during historical mining operations and more recent exploration activities. The Property is connected to the Moroccan national electricity grid, which currently supplies power to the exploration camp and associated facilities. Water for exploration activities is obtained from historical underground mine workings and groundwater wells located within the Property.

Existing exploration infrastructure includes administrative offices, drill core logging and storage facilities, the Afrilab sample preparation laboratory, drill contractor workshops, drill contractor accommodation camps and associated storage areas. These facilities support ongoing drilling, geological logging, sampling and other exploration-related studies.

Numerous historical mine access roads, tracks, shafts and underground workings occur across the Property, providing access to the principal historically identified mineralized zones and the historical underground mine (Figure 5‑1). The Project is located approximately 16 km from Tinejdad, which provides basic commercial services, accommodation and supplies, while the regional centre of Errachidia, approximately 70 km northeast of the Project, provides a broader range of commercial services, equipment suppliers, medical facilities and an international airport. Specialist equipment and supplies can be sourced from the larger commercial centres of Marrakech and Casablanca.

SRK considers that the existing infrastructure is adequate to support the current exploration activities being undertaken by Aya.

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Figure 5‑1    View looking across the Boumadine project area

5.3Climate

The Project lies within the eastern Anti-Atlas Mountains and experiences a hot desert (BWh) climate under the Köppen climate classification. The High Atlas Mountains form an effective climatic barrier to the north, limiting the influence of Mediterranean weather systems resulting in an arid, Saharan climate.

Summers are typically hot and dry, with maximum daytime temperatures commonly exceeding 40°C, particularly during July and August. Winters are generally mild, with daytime temperatures remaining moderate, although night-time temperatures can fall to near freezing during the coldest months. Average annual rainfall is low and is concentrated primarily during the winter months. Humidity is generally low throughout the year, while sunshine duration remains consistently high. Climatic records from the nearby town of Tinejdad indicate average summer maximum temperatures of approximately 38°C, winter average temperatures between 9°C and 15°C, and annual precipitation typically less than 110 mm.

Seasonal weather conditions are not considered to impose any significant constraints on exploration or future mining operations.

5.4Physiography

The Project is situated within the northwestern portion of the Ougnat Massif, where the landscape is characterised by rugged desert terrain comprising low hills, ridges and broad intervening valleys. Elevations across the Property range from approximately 980 m to 1,300 m above sea level, with local relief being typically controlled by the volcanic and volcano-sedimentary rocks that host the Boumadine mineralization.

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The landscape exhibits characteristics typical of the eastern Anti-Atlas, with extensive areas of exposed bedrock, thin soils and limited vegetation cover (Figure 5‑2 and Figure 5‑1). Drainage is predominantly ephemeral, with surface water generally restricted to seasonal watercourses following rainfall events.

Vegetation is sparse and consists primarily of drought-resistant desert species including Acacia raddiana and Tamarix amplexicaulis, together with scattered grasses adapted to arid conditions. More densely vegetated areas occur locally within oasis environments associated with permanent water sources, where date palm cultivation and small-scale agriculture are practised.

Figure 5‑2 shows a view looking across the historic Boumadine mine infrastructure, demonstrating the typical relief and vegetation of the area., whilst Figure 5‑1 also shows the access tracks used by Aya and its contractors and consultants to access the Project area.

Figure 5‑2    View looking over historical Boumadine mine

5.5Local Resources

The principal activity within the region is agricultural, focussed in river valleys and a number of oasis settlements, including Tinejdad. Date palm cultivation is the dominant agricultural activity, although smaller areas of mixed farming also occur where water resources are available.

The local population includes both permanent settlements and traditional Amazigh (Berber) nomadic communities, whose livelihoods are based principally on camel and goat husbandry. Seasonal movement of livestock through the Project area is common and represents a longstanding land use within the region.

The nearest major service centre is Errachidia, which provides commercial services, accommodation, medical facilities, equipment suppliers and an international airport. Additional services and accommodation are available in Tinejdad, whilst specialist equipment and supplies can be sourced from the larger commercial centres of Rabat, Casablanca and Marrakech. A local workforce suitable for exploration activities is available from the surrounding communities, with additional skilled personnel readily obtainable from elsewhere within Morocco.

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Overall, SRK considers that the availability of local infrastructure, services, utilities and labour is adequate to support the current exploration activities and potential future mining activities at Boumadine.

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6.0HISTORY

6.1Exploration and Mining History

The Boumadine Project has a long history of mineral exploration and small-scale mining (both artisanal and mechanised), with evidence of mineral extraction dating back several centuries. Historical and more modern surface workings are distributed along the principal mineralized structures and targeted both barite and polymetallic mineralization. The surface barite mining relates to modern mining activities, which are undertaken on a small scale for the collection of hand specimens and is ongoing within the Project area. Since the mid-twentieth century, the Project has been explored by several government organisations and private companies using geological mapping, geophysical surveys and drilling, progressively improving the geological understanding of the Project. The Boumadine deposit was also exploited using underground mining methods between 1986 and 1992 by BRPM.

6.2Antiquity to the sixteenth Century

Historical records indicate that Boumadine is located within one of the oldest known mining districts in Morocco. Mineralization is believed to have been exploited during antiquity, with more extensive mining activities undertaken during the fifteenth and sixteenth centuries, when Portuguese miners extracted high-grade oxidised (sulphide) mineralization exposed at surface. Numerous shallow pits, trenches and small underground workings remain visible throughout the Project area and extend to depths up to 20 m below surface. These historical workings occur along more than 4 km of the principal mineralized trend and provided the earliest indication of the continuity of the polymetallic vein system.

6.3BRPM Exploration (1956-1985)

Systematic exploration of the Project area commenced following the establishment of the Bureau de Recherches et de Participations Minières (BRPM), the former Moroccan Government mining agency responsible for national mineral exploration. Between 1956 and 1985, BRPM completed several phases of exploration designed to evaluate the economic potential of the known mineralized structures.

Early exploration activities included regional and detailed geological mapping, trenching, surface sampling and ground geophysical surveys, followed by diamond drilling to test the continuity of the mineralization at depth. These programs confirmed the presence of multiple steeply dipping polymetallic sulphide veins/domains containing Ag, Au, Pb, Zn and Cu mineralization.

During the 1970s and early 1980s, BRPM expanded exploration through additional drilling, underground development, mineralogical investigations and preliminary metallurgical testwork. Geological interpretations developed during this period established some of structural controls on mineralization and demonstrated the continuity of the principal mineralized zones along strike and down dip.

6.4Underground Development and Mining (1986–1992)

Following encouraging exploration results, underground development commenced in 1986 under the management of the Moroccan mining company SODIM. Mining was principally conducted as a pilot-scale operation to evaluate the continuity of the mineralized structures, metallurgical performance, and the economic potential of the deposit, rather than as a sustained commercial mining operation. Between July

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1986 and June 1988, SODIM treated approximately 23,700 t of mineralized material sourced from the -50 m level of the Central Zone, with reported average feed grades of 1.38% Pb and 6.0% Zn. Au and Ag recoveries were reported to be less than 30%, reflecting the limitations of the pilot flotation circuit employed at the time.

In June 1988, BRPM resumed control of the Project to undertake further industrial-scale metallurgical testwork. Between June 1988 and May 1989, BRPM processed a further 42,785 t of mineralized material from the -50 m level of the Central Zone, grading 1.57% Pb and 4.97% Zn, through the existing flotation plant. Although metallurgical performance improved, precious metal recoveries remained relatively low.

In 1989, BRPM transferred (loaned) operation of the Project to the "Société de Développement du Cuivre de l'Anti-Atlas" (SODECAT), which continued underground development and pilot-scale mining. Mining activities were concentrated principally on the -100 m level of the Central Zone and the -120 m level of the South Zone, with limited mining also undertaken in the North Zone. During this period, approximately 1,570 m of surface diamond drilling, 111 m of shaft sinking, 1,058 m of underground development and 187 m of raises were completed, together with the rehabilitation and extension of the existing underground infrastructure. In parallel with mining operations, underground drilling, channel sampling, metallurgical investigations and engineering studies were undertaken to further evaluate the deposit.

Mining operations ceased in 1992 after SODECAT concluded that the operation was only marginally profitable under the prevailing economic conditions. Despite the relatively limited scale of historical mining, the underground development and associated geological information have provided insights into the geometry, continuity and controls on the mineralized structures. At the effective date of this report, the majority of the historical underground workings are flooded and are currently utilized by the Company as a source of water for exploration activities.

6.5BRPM Exploration (1993–1998)

Following the cessation of mining, BRPM resumed exploration activities to reassess the remaining mineral potential of the Property. Work completed during this period included additional diamond drilling, geophysical surveys and a reassessment of the geological interpretation. These programs culminated in a MRE in 1998 (not compliant with modern CRIRSCO reporting codes), which represented the most comprehensive assessment of the deposit completed prior to the involvement of Maya.

Further details of the historical MREs are presented in Section 6.8.

6.6Maya Gold & Silver / Aya Gold & Silver (2013–Present)

On 9 October 2012, Maya entered into a joint venture agreement with the Office National des Hydrocarbures et des Mines (ONHYM) for the acquisition, development and exploitation of the Boumadine deposit. Exploration activities recommenced in 2013 following the establishment of Boumadine Global Mining SA, the operating company responsible for the Project.

Initial work programs focused on validating historical exploration results through geological mapping, surface prospecting and systematic grab sampling of the historical mineralized structures. These early investigations confirmed the presence of extensive polymetallic mineralization and identified numerous exploration targets beyond the historical mine areas.

Diamond drilling (DD) recommenced in 2017 with a program designed to confirm the historical geological interpretation and Mineral Resource. Between 2017 and 2020, the Company completed approximately 12,700 m of diamond drilling targeting the South, Central, North (includes the Tizi and Imarirren) Zones. Additional exploration work was also undertaken, including drone topographic surveys, sampling of historical tailings,

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geological mapping and continued metallurgical investigations. Collectively, these programs formed the basis of the first CRIRSCO-compliant MRE completed for the Project and reported in accordance with CIM Code guidelines.

On 31 July 2020, Maya changed its name to Aya and exploration activities were significantly expanded. Exploration completed between 2022 and the effective date of this MRE included DD, reverse circulation (RC) drilling, airborne electromagnetic, magnetic and radiometric surveys, satellite-based hyperspectral surveys, geological mapping, trenching, systematic surface sampling alongside a structural study. These programs substantially improved the geological understanding of the deposit, extended known mineralization along strike and down dip depth, and identified several new exploration targets throughout the Project area.

The progressive increase in drilling density and geological understanding has supported successive MRE updates and is a principal reason for the changes reflected in the updated 2026 MRE presented in this report.

During this time, hydrogeological, geotechnical, mineral processing, environmental, social, and mining studies have all been initiated by Aya to support the previous and forthcoming Preliminary Economic Assessments (PEAs).

6.7Historical Production

Historical underground mining undertaken between 1986 and 1992 produced approximately 261,485 t of mineralized material from four underground mining levels. According to BRPM (1998), production comprised 22,355 t from the Central Zone (-50 m level) between 1986 and 1987, 48,130 t from the Central Zone (-50 m level) between 1988 and 1989, 132,000 t from the Central Zone (-50 m and -100 m levels) between 1989 and 1992, and 59,000 t from the South Zone (-70 m and -120 m levels) between 1989 and 1992, at reported average grades of 3.69 g/t Au, 328 g/t Ag and 0.94% Pb.

6.8Previous Mineral Resource Estimates

Four previous MRE’s have been completed for the Boumadine Project. These comprise the historical MRE completed by the BRPM in 1998 (Table 6‑1), this estimate was not compliant with modern CRIRSCO reporting codes. The three other MRE’s were NI43-101 compliant estimates.

The NI43-101 compliant MRE’s were completed by GoldMinds Geoservices Inc in May 2019 (Table 6‑2), P&E Mining Consultants Inc. in May 2024 (Table 6‑3) and by Aya in March 2025 (Table 6‑4). The Aya estimate, the 2025 MRE, and model was used as the basis of a Preliminary Economic Assessment (PEA) dated December 2025.

All previous MRE and Mineral Resource Statements have been superseded by the SRK grade estimation, the 2026 MRE, presented in this Technical Report.

Table 6‑1    BRPM 1998 Boumadine MRE (not compliant with modern CRIRSCO reporting codes)

Category			Tonnes			Au (g/t)			Ag (g/t)			Zn (%)			Pb (%)
Certain			280,775			3.17			171			3.6			0.7
Very Probable			267,670			2.27			155			3.27			1.02
Probable			231,230			1.06			175			3.03			0.86
Possible			450,000			1.52			95			3.51			1.05

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Table 6‑2    GoldMinds Geoservices Inc 2019 Boumadine MRE (effective date 15 February 2019)

Classification			Tonnes			Au (g/t)			Ag (g/t)			Zn (%)			Pb (%)			Ge (g/t)			USDEq			Au eq (g/t)			Au eq (oz)
Total Measured			337 000			3.89			142.12			0.54			0.13			0			251.65			6.02			65,000
Total Indicated			2,195,000			1.57			127.88			3.06			1.2			3.9			250.45			6			423,000
Total Meas & Ind			2,532,000			1.88			129.77			2.73			1.06			3.38			250.61			6			489,000
Total Inferred			6,451,000			1.56			89.55			2.73			1.16			5.08			222.96			5.31			1,102,000

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Table 6‑3    P&E Mining Consultants Inc. 2024 Boumadine MRE (effective date 23 February 2024)

Class			Cut-off NSR US$/t			Tonnes (kt)			Average Grade																					Contained Metal
									Ag (g/t)			Au (g/t)			Cu %			Pb %			Zn %			AgEq (g/t)			AuEq (g/t)			Ag (koz)			Au (koz)			Cu (kt)			Pb (kt)			Zn (kt)			AgEq (koz)			AuEq (koz)
Pit-Constrained
Indicated			95			995			145			2.71			0.17			1.58			5.92			634			7.4			4,647			87			2			16			59			20,299			237
Inferred			95			8,474			103			2.97			0.1			0.79			2.54			475			5.54			28,087			808			8			67			215			129,478			1,510
Out-of-Pit
Indicated			125			1,046			82			2.33			0.11			0.58			2.79			402			4.69			2,751			78			1			6			29			13,533			158
Inferred			125			15,096			76			2.42			0.07			0.87			2.19			389			4.53			36,653			1,175			11			131			330			188,663			2,198
Total
Indicated			95/125			2,042			113			2.51			0.14			1.07			4.32			515			6.01			7,398			165			3			22			88			33,832			395
Inferred			95/125			23,569			85			2.62			0.08			0.84			2.32			420			4.89			64,740			1,983			19			198			546			318,141			3,708

Table 6‑4    Aya Gold & Silver Inc. 2025 Boumadine MRE (effective date 24 February 2025)

Class			Cut-off NSR US$/t			Tonnes (kt)						Average Grade																		Contained Metal
									Ag (g/t)			Au (g/t)			Cu %			Pb %			Zn %			AgEq (g/t)			AuEq (g/t)			Ag (koz)			Au (koz)			Cu (kt)			Pb (kt)			Zn (kt)			AgEq (koz)			AuEq (koz)
Pit-Constrained
Indicated			95			3,920			94.3			2.99			0.13			0.84			2.95			476.5			5.3			11,881			343			5.1			33			116			60,051			667
Inferred			95			14,258			89.7			2.89			0.1			0.81			2.38			450			5			41,135			1102			14.3			115			339			206,293			2,293
Out-of-Pit
Indicated			125			1,249			80.1			2.11			0.08			0.87			2.32			358.2			3.98			3,216			106			1			11			29			14,382			160
Inferred			125			14,938			74.3			2.39			0.07			0.82			1.85			356.9			3.97			35,669			1,294			10.5			122			276			171,393			1,905
Total
Indicated			95/125			5,169			90.8			2.78			0.12			0.85			2.8			447.9			4.981			15,097			449			6.1			44			145			74,433			827
Inferred			95/125			29,196			81.8			2.63			0.08			0.82			2.11			402.4			4.473			76,804			2,396			24.8			237			615			377,686			4,198

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7.0GEOLOGICAL SETTING AND MINERALIZATION

7.1Regional Geology

Boumadine is located within the eastern Anti-Atlas Mountains of southeastern Morocco, on the northwestern margin of the Ougnat Massif (also referred to as the Ougnat Boutonnière). The Anti-Atlas forms a major east-northeast-trending mountain belt extending for more than 600 km across southern Morocco and comprises a series of Neoproterozoic inliers (boutonnières) surrounded by younger Palaeozoic sedimentary rocks. The region has exposed rocks associated with the Pan-African Orogeny (approximately 900–550 Ma), which record multiple phases of volcanism, sedimentation, deformation and granitoid intrusion.

Within the Ougnat Massif, a Neoproterozoic metasedimentary basement is unconformably overlain by volcano-sedimentary rocks of the Ouarzazate Supergroup. These rocks comprise felsic to intermediate volcanic and volcaniclastic units together with subordinate mafic volcanic rocks and sedimentary horizons. The Ougnat Massif was subjected to Neoproterozoic shearing, which generated regional-scale faults, which trend northeast – southwest (030/210°). Subsequent Hercynian deformation resulted in the reactivation of these earlier Pan-African structures, producing a complex structural framework that exerted a fundamental control on the emplacement of hydrothermal mineralization throughout the Anti-Atlas.

Recent regional geological studies by Gaboury (2022, 2023) further suggest that the Ougnat Massif may represent part of a large Neoproterozoic volcanic graben or caldera complex, within which the Boumadine deposit is situated (Figure 7‑1). This interpretation is based on the distribution of volcanic and volcano-sedimentary rocks, felsic volcanic centres and regional structural architecture, and provides a possible tectono-magmatic framework for the development of the extensive hydrothermal system associated with the Boumadine deposit.

The volcanic and volcano-sedimentary rocks have been grouped into three formations:

•Tamerzaga-Timrachine Formation ("TTF"): Consists of a volcano-sedimentary sequence comprising felsic crystal tuffs, rhyodacitic ignimbrites, volcaniclastic rocks and subordinate mafic tuffs.

•Isilf-Ouinou-Oufroukh Formation ("IOF"): Consists of volcano-sedimentary rocks, including tuffs and breccias, andesitic flows, and fine- to coarse-grained sedimentary rocks.

•Aoujane-Aissa-Akchouf Formation ("AAA"): Consists of ignimbrites, dacite domes and flows, and andesitic flows.

Boumadine is hosted within the TTF, which comprises felsic volcanic and volcaniclastic rocks overlying mafic volcanic units. Subsequent geological evolution included emplacement of felsic subvolcanic intrusions together with later mafic dykes, which locally offset or truncate the mineralized system. The volcanic succession has undergone extensive hydrothermal alteration associated with polymetallic Au-Ag-Zn-Pb mineralization.

Similar Neoproterozoic volcanic-hosted precious and base metal deposits occur elsewhere throughout the Anti-Atlas Province, reflecting the regional-scale metallogenic significance of the Pan-African tectonic and magmatic events.

Figure 7-1 is a regional geology of the Anti-Atlas and the Ougnat Massif.

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Figure 7‑1    Regional geology of the Anti-Atlas and the Ougnat Massif. (A) Geological map showing the distribution of the Neoproterozoic inliers (boutonnières) within the Anti-Atlas (after Bouabdellah and Levresse, 2016). (B) Simplified geological map of the regional geological setting (after Duplessis et al., 2019). The red dashed oval outlines the interpreted graben or caldera. Figure reproduced from Gaboury (2022).

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7.2Property Geology

As noted above, the mineralization is hosted by Neoproterozoic volcanic and volcano-sedimentary rocks of the TTF, comprising both mafic and felsic units. The host succession comprises felsic volcanic and volcaniclastic rocks that unconformably overlie mafic volcanic rocks. The felsic succession consists principally of crystal-rich felsic tuffs and volcaniclastic units containing quartz-eye fragments and plagioclase phenocrysts, whereas the underlying mafic succession comprises crystal-rich mafic tuffs containing abundant plagioclase and subordinate amphibole phenocrysts. Locally, the mafic volcanic rocks contain sedimentary clasts, supporting their interpretation as volcaniclastic deposits emplaced in a subaqueous volcanic environment.

The distinction between the felsic and mafic tuffs is an important component of the geological interpretation and appears to influence the distribution and intensity of mineralization; however, this distinction is locally difficult to recognize or is absent from the historical drillhole logging. The principal mineralized corridors are interpreted to be predominantly structurally controlled and extend across the different volcanic and intrusive lithologies.

The geology varies between the principal mineralized zones, as shown in Figure 7‑2. The South and Central zones comprise a combination of felsic volcanic and volcaniclastic rocks, felsic subvolcanic intrusions and mafic tuffs, with mineralization hosted within both the felsic and mafic successions. In contrast, mineralization in the northern area, particularly at Tizi and Imarirren, is more strongly associated with the mafic volcanic succession.

The felsic volcanic and volcaniclastic succession comprises a complex assemblage of felsic volcanic rocks, hydrothermal breccias and felsic subvolcanic intrusions that have been affected by multiple stages of faulting, hydrothermal alteration and sulphide mineralization. Mineralization occurs within predominantly structurally controlled corridors. The Imarirren and Tizi zones, grouped within the North Zone, are located west of the principal Boumadine trend associated with the Central and South zones. Rhyolitic intrusions and volcanic breccias occur locally across the Project. Geological studies undertaken by Gaboury indicate that the stratigraphy and intrusive relationships across the Property are more complex than previously recognized, with repetition of the volcanic sequence interpreted to result locally from synvolcanic faulting.

Drilling completed by Aya has demonstrated that the volcanic stratigraphy includes numerous felsic subvolcanic intrusions occurring as dykes, sills and dome-like bodies. These intrusions display massive, flow-banded, autobrecciated and volcaniclastic textures, indicating emplacement within a shallow volcanic environment. The felsic intrusions are spatially associated with the principal mineralized corridors and are interpreted to have locally provided favourable permeability for hydrothermal fluid flow. Later mafic dykes, including locally porphyritic varieties, cross-cut the volcanic stratigraphy and mineralized structures and are interpreted to post-date the principal mineralizing event. These dykes are commonly north-south oriented, although their orientation varies locally, and they are more abundant in the northern area of the deposit.

The mafic dykes are compositionally similar to the mafic tuffs, contain abundant plagioclase phenocrysts, and occur both as sills within the felsic tuffs and as dykes cutting felsic intrusive rocks. Gaboury (2022, 2023) interpreted these relationships to suggest that at least part of the mafic magmatism was broadly synchronous with the felsic magmatism, reflecting the evolution of a bimodal volcanic system.

Hydrothermal alteration is widespread throughout the Property and is closely associated with the mineralized structures. The felsic volcanic sequence is characterised by pervasive phyllic alteration comprising quartz-sericite-pyrite assemblages that form broad alteration halos surrounding the mineralized zones. Localized zones of kaolinite, pyrophyllite and silicification occur adjacent to mineralization and are interpreted to reflect the circulation of relatively acidic hydrothermal fluids. The underlying mafic volcanic rocks exhibit predominantly

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propylitic alteration characterised by chlorite and epidote, with increasing chloritization and pyrite developed proximal to mineralized structures. Black chlorite alteration has also been recognized immediately adjacent to some massive sulphide zones and is interpreted to represent a proximal hydrothermal alteration assemblage.

Extensive near-surface weathering has produced an iron-rich oxidised profile ("chapeau de fer") developed above the sulphide mineralization. The depth of oxidation is variable but typically ranges from 10 m to 50 m below surface, although local variations occur depending on topography and structural setting. The oxidised zone is dominated by goethite and jarosite with subordinate hematite and reflects prolonged supergene oxidation of the sulphide-bearing vein systems.

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Figure 7‑2    Simplified geological map of the Boumadine Project showing the principal lithological units, intrusive rocks, major structures and the distribution of the North (Imarirren and Tizi), Central, South, mineralized zones (modified after Bouabdellah and Levresse, 2016; reproduced by Gaboury, 2022).

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7.3Structural Geology

The structural framework of the Boumadine Project has been described by several previous studies, including BRPM (1998), GoldMinds Geoservices (2019) and P&E Mining Consultants (2024). Earlier interpretations considered the deposit to comprise a series of steeply dipping, structurally controlled polymetallic veins developed within regional Pan-African structures that were subsequently affected by later deformation. Regional-scale faults, together with associated subsidiary fractures and structures, were interpreted to have acted as conduits for hydrothermal fluid flow and mineral deposition.

The principal mineralized structures generally strike northwest-southeast to north-northwest-south-southeast (340°–350°/160°–170) and dip steeply towards either the northeast or southwest. The principal mineralized structures are cross-cut and locally offset by later northeast-striking (020-200°) steeply dipping faults. The northeast striking faults with the larger offset have been used to divide the deposit into its three distinct areas (South, Central and North). Mafic and felsic dykes also cross-cut the volcanic succession and occur in several orientations; earlier andesite dykes are reported to strike approximately 160-340, whereas later dyke swarms commonly strike approximately north-south (0-180) to north-northeast (030-210°). Late mafic dykes are particularly abundant within the Tizi and Imarirren areas. These interpretations formed the basis of the geological models supporting the 2019 and 2024 MREs.

In addition to this, geological investigations undertaken by Aya, including detailed drill core logging and geological studies completed by Gaboury (2022, 2023), refined the structural interpretation of the deposit (Figure 7‑3). These studies indicate that the principal mineralized corridors are spatially associated with steeply dipping synvolcanic extensional fault systems that accommodated vertical displacement during deposition of the volcanic succession. Repetition of the felsic and mafic volcanic stratigraphy across several of the principal mineralized corridors provides support for this interpretation. Gaboury interpreted these structures as long-lived synvolcanic faults that focused hydrothermal fluid flow and controlled the emplacement of polymetallic sulphide mineralization.

Interpretation of Worldview-3 alteration imagery (Figure 7‑4), geological mapping and drilling indicates that the principal mineralized structures are spatially associated with major structural corridors and felsic subvolcanic intrusions. Gaboury (2022, 2023) proposed that mineralization developed preferentially within structurally favourable zones associated with synvolcanic faulting, including dilational zones and fault intersections. Hydrothermal breccias, massive to semi-massive sulphides, pyrite stockworks and felsic subvolcanic intrusions are spatially associated with these structures, with repeated fault activity providing pathways for hydrothermal fluid circulation and mineral deposition. Felsic subvolcanic intrusions are also interpreted to have locally provided favourable permeability and, in some areas, an additional magmatic fluid contribution to the hydrothermal system.

Gaboury also identified cross-cutting fault sets that locally intersect the principal northwest- to north-northwest-trending mineralized corridors. In particular, Gaboury (2023) identified approximately 020-200°-trending structures and interpreted intersections between these structures and the principal mineralized trend as possible controls on local thickening and metal enrichment.

The current geological interpretation adopted for the 2026 MRE incorporates these updated structural concepts while maintaining continuity with the previous geological framework.

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Figure 7‑3    Conceptual structural model proposed by Gaboury (2022) illustrating the interpreted relationship between synvolcanic normal faulting, dextral strike-slip faulting, felsic subvolcanic intrusions, pyrite stockwork and the development of semi-massive sulphide mineralization at the Boumadine deposit (after Gaboury, 2022).

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Figure 7‑4    WorldView-3 alteration map and structural interpretation presented by Gaboury, 2022.

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7.4Mineralization

The polymetallic sulphide mineralization at the Boumadine deposit generally comprises 1 to 5 m wide, massive to semi-massive sulphide veins and lenses, although locally thicker zones exceeding 10 m are observed. The principal mineralized structures typically strike northwest-southeast (approximately 135°–315°) or, north-south (approximately 000°–180°), and generally dip steeply (>70°), predominantly towards the northeast. These structures typically host the highest-grade mineralization within the Project. The massive sulphide (>70% logged visible sulphide) and semi-massive sulphide (>30% and <70% logged visible sulphide) mineralization is composed predominantly of pyrite, with lesser amounts of arsenopyrite, sphalerite, galena and traces of chalcopyrite. Previous publications have also noted minor amounts of cassiterite, silver-rich sulphosalts, stannite, enargite, bismuthinite, native copper and bismuth.

The massive to semi-massive sulphide mineralization is generally surrounded by a 1 to 10 m thick, and locally up to 20 m thick, halo containing approximately 10 to 30% disseminated sulphides, predominantly pyrite, although the sulphide content within the halo can be considerably lower.

Quartz-carbonate-galena-sphalerite veins are also observed across the Project area and can be laterally continuous at surface. These veins have locally been exploited for barite, principally for the collection of hand specimens.

Gaboury (2022, 2023) suggested that the mineralization can be broadly characterised by four recurring styles that record the evolution of the Boumadine hydrothermal system: (i) massive to semi-massive sulphide veins developed within quartz-carbonate veins and fault zones; (ii) sulphide-rich hydrothermal breccias formed within zones of brittle deformation; (iii) replacement (impregnation) sulphide mineralization developed within porous felsic volcanic and volcaniclastic rocks; and (iv) pyrite-dominated stockwork veinlets and disseminated sulphides developed within alteration halos surrounding the principal mineralized structures. These styles commonly occur together within the principal mineralized corridors and reflect multiple episodes of fault movement, hydrothermal fluid flow, brecciation and sulphide deposition. Gaboury (2022) developed a conceptual model illustrating how these mineralization styles may occur in different configurations depending on the interpreted fluid pressure and structural setting (Figure 7‑6). Under suprahydrostatic pressure conditions, semi-massive sulphide veins may be surrounded by pyrite stockworks forming a damage zone or by pyrite veinlets developed approximately parallel to the principal sulphide vein. Under hydrostatic pressure conditions, mineralization may occur within synvolcanic fault-related hydrothermal breccias (BX3), characterised by silicification, disseminated pyrite, rounded pyrite fragments and angular lithic fragments. Semi-massive sulphide mineralization may be absent from these breccias or may occur as impregnation developed asymmetrically along one wall or symmetrically along both walls of the mineralized breccia zone. The orientation of the smaller pyrite veinlets can be highly variable.

Representative mineralization textures are shown in Figure 7‑6. Gaboury also noted that the orientation of the smaller pyrite veinlets can be highly variable.

Geochemical relationships identified during previous studies indicate a strong positive correlation between Au, Ag and Cu, together with a weaker positive correlation between Zn, Pb and molybdenum (Mo). More recent mineralogical investigations undertaken by Aya indicate that approximately 95% of the Au is associated with pyrite and arsenopyrite and occurs in microscopic to submicroscopic form. Gaboury (2023) further recognized that local enrichment in Au, Ag, Zn and Pb is associated with thermal refinement of massive sulphides, while elevated Au grades within pyrite stockworks are closely associated with arsenopyrite development.

Petrographic observations from mineralized samples at Boumadine indicate that sphalerite locally replaces earlier pyrite and is subsequently followed by galena, reflecting multiple stages of hydrothermal mineral

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deposition. U-Pb single-zircon dating of a “chonolithic” rhyolite intrusion that cross-cuts the mineralized veins yielded an age of 553 ± 16 Ma (Levresse, 2001), which has been interpreted to provide a late Neoproterozoic maximum age for the mineralization.

Figure 7‑5    Examples of sulphide mineralization textures from the Boumadine Project. A) Brecciated and slightly oxidized pyrite-rich mineralization with quartz veinlets from the Central Zone muck pile. B) Pyrite-sphalerite mineralized material, sample BOU2012-02, collected from the muck pile near Shaft A of the Central Zone. C) Galena-rich mineralization with pyrite, pyrrhotite, chalcopyrite, sphalerite and later quartz veining from the Central Zone muck pile. D) Reflected-light microphotograph showing pyrite with sphalerite and galena Modified after Boily (2014)

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Figure 7‑6    Conceptual model illustrating five configurations of sulphide mineralization at Boumadine under interpreted suprahydrostatic and hydrostatic pressure conditions, including semi-massive sulphide (SM) veins, pyrite (Py) stockworks and veinlets, and synvolcanic fault-related hydrothermal breccias (BX3) with variable development of semi-massive sulphide mineralization (after Gaboury, 2022)

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8.0DEPOSIT TYPES

Earlier studies (prior to Gaboury involvement) interpreted mineralization to be largely confined to altered felsic volcanicand volcaniclastic rocks, as mineralized outcrops are exposed almost exclusively within felsic tuffs. More recent drilling completed by Aya has demonstrated that significant mineralization also extends into the underlying mafic volcanic sequence, in some cases the mineralized structures appear to be locally more continuous, thicker and of higher grade. This observation has led Gaboury (2022, 2023) to propose that the mafic volcanic succession represents an important host to the mineralizing system and may reflect a deeper, more confined part of the hydrothermal system.

Gaboury (2022, 2023) further proposed that Boumadine formed within a shallow submarine volcanic environment developed in an extensional basin or graben setting, rather than in a purely subaerial epithermal environment. Under this interpretation, circulation of low-temperature saline convective fluids produced widespread propylitic alteration, while episodic input of acidic magmatic fluids associated with subvolcanic felsic intrusions generated local quartz-sericite-pyrite and advanced argillic alteration together with Au-Ag mineralization. Mixing of these fluid types is interpreted to have produced the observed polymetallic sulphide assemblage dominated by pyrite, and lesser amounts of sphalerite and galena, with subordinate chalcopyrite and arsenopyrite, thereby explaining the transitional characteristics between epithermal and VMS-style mineralization.

This interpretation also provides a geological explanation for the laterally extensive phyllic alteration within the felsic volcanic sequence, the continuity of the mineralized structures into the mafic volcanic succession, and the coexistence of vein-hosted, replacement and massive sulphide mineralization within the same hydrothermal system. Nevertheless, the precise genetic classification of the Boumadine deposit remains the subject of ongoing investigation. The 2026 MRE is based on the observed geological, structural and mineralization characteristics of the deposit and does not depend on any particular genetic interpretation.

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9.0EXPLORATION

This section summarises exploration work completed by Maya and Aya at the Boumadine Project. Historical exploration completed prior to Maya’s acquisition of the Project is summarised in Section 6 above.

Since acquiring the Project in 2012, Maya and Aya have undertaken systematic exploration comprising geological mapping, specialist geological studies, geochemical sampling, airborne geophysical surveys, hyperspectral surveys, grab sampling, trenching, tailings sampling and extensive drilling programs. Exploration has progressively expanded from confirmation of the historical Boumadine deposit to regional exploration across the wider property, together with infill drilling undertaken to support the 2026 MRE.

As of 16 April 2026, the exploration database comprised 1,859 exploration records, including surface diamond drilling (DD), reverse circulation (RC) drilling, RC pre-collars with DD tails (MP), historical underground channel samples, underground production drillholes, and dedicated tailings drilling and sampling records. The database contains approximately 425,919 m of drilling and sampling completed by BRPM, Maya and Aya, together with historical underground exploration data (Table 9‑1), as well as drilling conducted in neighbouring Aya licences (Figure 9‑1). It should be noted that the latest assays associated with the database were issued from Afrilabs on the 28 February 2026.

The tailings deposit, and the shallow holes drilled to investigate it, associated with the historical underground mine has not been considered as part of this study.

Table 9‑1    Summary of Aya Database provided by Aya to SRK

Company

Year(s)

Type (Code)

Exploration Record Type

Number of Records

Total Metres (m)

BRPM

1963–1985

DD

Surface diamond drillholes

118

15,590.81

BRPM (SODIM & SODECAT?)

1964–1966

PA

Underground percussion (production) drillholes

18

391.03

BRPM (SODIM & SODECAT?)

1964–1992

CS

Underground face channel samples

172

680.3

BRPM (SODIM & SODECAT?)

1,964

UG-Channel

Underground wall channel samples

4

789

BRPM

Historical*

SS

Surface grab/channel sampling

207

351.26

BRPM Subtotal

519

17,802.40

Maya

2017–2020

DD

Surface diamond drillholes

87

12,158.71

Maya

2,018

TA

Diamond drillholes completed on historical tailings

49

187.28

Maya Subtotal

136

12,345.99

Aya

2022–2026

DD

Surface diamond drillholes

1,032

349,968.29

Aya

2025

DD†

Surface diamond drillholes completed on historical tailings facilities

160

673.5

Aya

2023–2025

RC

Reverse circulation drillholes

91

18,488.90

Aya

2024–2025

MP (RC/DD)

Reverse circulation pre-collar with diamond tail

81

27,313.20

Aya Subtotal

1,204

395,770.39

Grand Total

1963–2026

1,859

425,918.78

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Figure 9‑1    Inclined image showing the exploration database provided by Aya plotted in relation to the Boumadine Mining Licence (black), collars coloured by Year with traces coloured grey

9.1Geochemical Sampling/Grab Samples

Surface geochemical sampling has comprised grab sampling, rock-chip sampling, which was completed during geological mapping and prospecting programs, the exploration database contains 207 historical surface grab/channel sampling records demoted as SS in Aya’s database. Sampling was undertaken to characterise exposed mineralization, evaluate new exploration targets and support geological interpretation. Historical tailings have also been systematically sampled to evaluate their precious- and base-metal content, though the associated assay data does not appear to be contained within the Aya database but is noted in the P&E 2024 MRE report.

These samples were used for exploration purposes only and were not used directly in the 2026 MRE.

9.2Historical Underground Sampling

Historical underground mining completed by BRPM generated a valuable geological dataset comprising 172 underground face channel samples, denoted as “CS” in Aya’s database; four underground wall channel samples, denoted as “UG-Channel”; and 18 underground percussion (production) drillholes, denoted as “PA”.

9.3Bulk Sample

No specific bulk samples have been taken by Aya or Maya at the Project to date, historically the underground mining was used to obtain samples for a pilot-scale operation to evaluate the continuity of the mineralized structures, metallurgical performance, and the economic potential of the deposit, rather than as a sustained commercial mining operation, see Section 6.4 for further details on this.

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9.4Hyperspectral Surveys

At the end of 2021, Compagnie Générale de Géophysique – Veritas (CGG) completed a satellite-based geological and hyperspectral mineral mapping study over the Boumadine Property on behalf of Aya (Figure 9‑2). The objective of the study was to identify areas exhibiting hydrothermal alteration and favourable structural features associated with the Boumadine mineralizing system, and to place these within a geological framework to support exploration targeting. The program generated geologically meaningful target areas that were subsequently verified during geological mapping campaigns completed in 2022 and 2023.

The study was completed in two stages. The first comprised a regional interpretation at a scale of approximately 1:25,000, aimed at defining the principal structural controls and regional alteration patterns across the property. The second consisted of detailed mapping at a scale of approximately 1:5,000 focused on the Boumadine Mining Licence, providing higher-resolution structural and alteration information over the known mineralized corridor(s).

Multiple satellite datasets were integrated during the interpretation. Sentinel-2 imagery was used as the primary geological base, with Copernicus digital elevation data incorporated to enhance lithological and structural interpretation. ASTER multispectral imagery was utilized to identify clay-, iron- and carbonate-rich alteration assemblages associated with hydrothermal systems and intrusive activity. For detailed mapping, high-resolution WorldView-3 imagery was employed for both structural interpretation and spectral mineral mapping.

The interpreted alteration mineral maps showed a strong spatial correlation with the principal mapped structures and known mineralized zones/corridors. Significant argillic alteration was identified along the northwest-trending fault in the northeastern part of the Boumadine Mining Licence, together with additional argillic alteration developed along a northeast-trending fault immediately west of the principal mineralized corridor. The hyperspectral mineral products also exhibited a close correspondence with the underlying geology, particularly within the North Zone where altered rhyodacite-rhyolite ignimbrite units were clearly delineated. These results provided additional confidence in the structural interpretation and were subsequently integrated with geological mapping, airborne geophysical data and drilling to prioritize follow-up exploration targets.

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Figure 9‑2    Map of the 2022 Hyperspectral Survey Completed on Boumadine, (source Aya 2025)

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9.5Geophysical Surveys

Historical geophysical surveys completed by BRPM included self-potential (SP), induced polarization (IP), magnetic, gravimetric and controlled source audio-frequency magnetotelluric (CSAMT) surveys, which assisted in delineating the principal mineralized structures and guiding historical drilling programs.

Aya significantly expanded the geophysical coverage of the Boumadine Property. In March 2022, Geotech Ltd. completed a helicopter-borne airborne geophysical survey covering approximately 366 line-km (33 km²) over the Boumadine Mining Licence. Traverse lines were flown east-west at 100 m spacing, with 1,000 m-spaced tie lines. The survey incorporated a Versatile Time Domain Electromagnetic (VTEM) survey (Figure 9‑3), a caesium vapour magnetometer, and an RSI ARGS RSX-5 radiometric spectrometer. The objective of the program was to acquire high-resolution magnetic, electromagnetic and radiometric datasets capable of mapping the known mineralized corridor and identifying potential extensions beneath shallow cover.

The 2022 survey successfully identified a prominent north-south elongate conductive (low-resistivity) anomaly measuring approximately 6 km by 1.5 km, which coincides closely with the historical Boumadine mine workings and the principal mineralized corridor. The conductive anomaly also spatially corresponds with the northern portion of a magnetic low, accurately delineating the known mineralization and supporting the continuation of the prospective structural corridor to both the north and south of the deposit extents known at the time. These datasets were subsequently integrated with geological mapping, hyperspectral interpretation and drilling to refine the geological model and prioritize exploration targets.

Following the success of the 2022 program, Aya commissioned Expert Geophysics Limited (EGL) to complete a second helicopter-borne MobileMT electromagnetic and magnetic survey between February and July 2024. The survey covered the Boumadine (contains the deposit itself), Boumadine West (area to the west of the mining licence) and Boumadine East (area to the east of the mining licence) blocks, encompassing all permits held by Aya at that time. The MobileMT system incorporated airborne three-component electromagnetic sensors together with a caesium vapour magnetometer to provide improved conductivity and magnetic imaging to depths exceeding those achieved during the 2022 survey.

Interpretation of the 2024 survey identified numerous conductive anomalies exhibiting orientations similar to the Boumadine deposit, including several previously unrecognized sub-parallel conductive corridors. Of particular significance was the identification of a large conductive anomaly located approximately 5 km west of the Boumadine deposit, together with extensions of the principal Boumadine conductivity anomaly to the south and additional northwest- and north-south-trending conductive features. These anomalies continue to provide priority targets for follow-up geological mapping and drilling.

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Figure 9‑3    2022 VTEM Magnetic Survey Completed at Boumadine (source Aya 2025)

9.6Geological and Structural Mapping

Geological mapping has been undertaken during serval major phases of exploration at the Boumadine Project. Between 2013 and 2016, Maya completed reconnaissance geological mapping and grab sampling of the historically known mineralized structures to verify historical geological interpretations and identify targets for the initial drilling programs. Following the commencement of Aya's exploration programs, detailed geological mapping was undertaken between 2022 and 2025, including specialist studies by Gaboury, to improve the understanding of the lithology, alteration, structural geology and mineralization. This work was integrated with airborne geophysical surveys, hyperspectral interpretation, drilling and geological modelling of the deposit.

Structural mapping identified two principal fault sets superimposed on the main northwest-southeast-trending mineralized corridor, which strikes approximately 340°–350°/160°–170°. The first fault set comprises

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northeast (approximately 030°) striking faults that intersect the principal mineralized structures and are interpreted to have influenced Au enrichment and a subsequent Zn mineralizing event. The second fault set comprises east-northeast (approximately 070°) striking faults that cross-cut both the main mineralized corridor as well as north-northeast (approximately 020°) trending structures. These faults are interpreted to represent a later structural event associated with Ag-Pb mineralization.

9.7Trenching

A single trench was completed by Aya during the 2022 exploration program was excavated between drillholes BOU-DD22-001 and BOU-DD22-006 on Section 8125N for a total of 37 m. No significant additional mineralization was identified, and subsequent exploration concentrated on drilling.

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10.0DRILLING

Drilling completed to date has comprised predominantly DD, undertaken to identify, delineate, extend and infill the Boumadine mineralized system and surrounding exploration targets located in adjacent licences. DD has also been completed on the historical tailings deposits, although these are not considered as part of the 2026 MRE. The drilling has provided the principal geological, structural, geotechnical and assay information supporting the geological interpretation and grade estimation associated with the 2026 MRE.

As of the effective date of the 2026 MRE, the exploration database contains 1,237 surface DD (377,717.81 m), 91 RC drillholes (18,488.90 m) and 81 MP drillholes (RC pre-collar with a diamond tail; 27,313.20 m), together with historical underground production drilling, underground channel sampling and surface sampling completed by previous operators, as summarised in Table 9‑1. In addition to the resource drilling programs, both Maya and Aya completed dedicated DD programs to investigate the historical tailings facilities, comprising 49 drillholes (187.28 m) completed by Maya during 2018 and 160 drillholes (673.50 m) completed by Aya during 2025. These tailings drillholes are presented separately in Table 9‑1 and are not included within the Mineral Resource drilling statistics in Section 14 below, as the tailings are not considered part of this study. It should be noted that information relating to the drillholes completed in 1992 by the SODECAT-BRPM were not able to be sourced by Aya and therefore are not contained within the Aya database.

Since 2022, Aya has significantly expanded the drilling program, resulting in an increase in the quantity of geological information available to support geological modelling and subsequent MRE updates. In 2023, Aya relogged and re-sampled all available historical DD core from DD completed between 2018 and 2020, comprising 77 drillholes totalling 9,510 m. Historical BRPM drill core and core from the 14 Maya DD completed during 2017, totalling 3,158.2 m, could not be located within Aya's core storage facilities and were therefore unavailable for relogging and re-sampling. These drillholes have been excluded from the 2026 MRE due to the lack of available drill core to support verification of the historical geological and assay data.

Maya and Aya have utilized two drilling contractors during the modern exploration programs: Geosond (Geosond) Maroc SARL and FTE Drilling. At the time of the SRK site visit in April 2026, Geosond and FTE Drilling were active on site, with each contractor operating five diamond drill rigs, Figure 10‑1 shows one of the DD drill rig set up for drilling at the Project. CT20 and CS140 drill rigs were used by (Geosond), whilst FTE used Versadrill and Marcotte drill rigs.

All RC drilling has been conducted by FTE drilling. The drill rigs are all aligned by Aya geologists prior to commencement of drilling.

Drillhole spacing is variable across the Project. However, the majority of the modern drilling used in the 2026 MRE was initially completed on nominal 100-200 m-spaced inclined drill fences, with subsequent infill drilling undertaken on approximately 50 m spacing. The infill drilling is ongoing, with data being collected to be incorporated into future studies. In certain locations, drillhole spacing is closer owing to variations in drillhole inclination or where additional drilling was planned to provide increased drill coverage.

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Figure 10‑1    DD drill rig set up at the Project during the SRK QP site visit

10.1Diamond Drilling

Surface diamond drilling forms the principal exploration dataset supporting the geological modelling and grade estimation. Historical drilling completed by BRPM between 1963 and 1985 comprised 118 surface DD totalling approximately 15,591 m. Maya was the first company to undertake modern surface DD at the Project, completing 87 drillholes between 2017 and 2020 for a total of approximately 12,159 m. Aya subsequently expanded the exploration program between 2022 and the effective date of this report, completing 1,032 surface DD drillholes totalling approximately 349,968 m.

Review of the assay database indicates that modern diamond drilling was completed using both HQ and NQ core diameters. Approximately 54% of the assayed diamond drill intervals were drilled using NQ core and 44% using HQ core. Both HQ and NQ core diameters have been utilized throughout the modern drilling programs completed between 2018 and 2026. Analysis of the drillhole database indicates that the majority of drillholes were initiated using HQ core before reducing to NQ at depth.

It should be noted that all drillholes are collared from surface, with the exception of BOU-DD24-455. According to Aya, BOU-DD24-455 is a daughter hole that was wedged from the parent drillhole BOU-DD24-352 at a depth of approximately 215 m below the collar of the parent hole. Consequently, the collar of BOU-DD24-455 does not coincide with the topographic surface.

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10.2Reverse Circulation Drilling

RC drilling was undertaken by FTE Drilling to improve drilling productivity. As of the effective date of the drilling database, Aya had completed 91 RC drillholes totalling 18,488.9 m. The RC drilling was undertaken using 6 m drill rods.

In addition, Aya completed 81 MP drillholes, comprising RC pre-collars followed by diamond drill tails, totalling 27,313.2 m. The MP drilling method was used where rapid penetration through overburden or barren rock was required, before continuing with diamond drilling to obtain geological, structural and geotechnical information within and around the mineralized intervals.

SRK’s QP did not observe an RC drill rig in operation during the site visit, as no RC drilling was being undertaken at that time.

10.3Collar Surveys

Historical drillhole collars completed by BRPM were surveyed using the survey methods available at the time of drilling; however, the procedures used for surveying these historical collars have not been documented. The historical surface drillhole collar coordinates were originally recorded in a local mine grid. As noted in Section 12.3, Maya commissioned GOULEX, in 2014 to compile and validate the available historical exploration and mining data. The scope of work included the verification and surveying of historical drillhole collars, shafts and tailings facilities using differential GPS (DGPS), with all surveyed coordinates recorded in the Project UTM coordinate system.

Modern drillhole collars completed by Maya and Aya were surveyed by a Moroccan company named ADETOP, who used a Trimble R8 station, which has an accuracy of 10 cm in the X, Y and Z direction or less. The surveyed collar coordinates were then incorporated into the Project database.

As part of the preparation of the 2026 MRE, SRK independently verified the locations of 14 drillhole collars during the site visit, as noted in Section 12.2 below.

10.4Downhole Surveys

Downhole deviation surveys were completed on all modern exploration drillholes to accurately define drillhole trajectories for geological interpretation and grade estimation. Aya provided the information in Table 10‑1 below, which notes the downhole survey instruments utilized during the drilling; the survey instrument is noted vary depending on the drilling contractor and year of drilling. Geosond initially employed Reflex EZ-Track instruments between 2022 and early 2023 drilling programs before introducing Devico DeviFlex instruments during 2023 and 2024. From 2024 onwards, Geosond progressively adopted Devico DeviGyro and subsequently SPT GyroMaster gyro survey instruments. FTE Drilling predominantly utilized Reflex EZ-Track instruments for drilling between 2024 to 2026, while both Reflex EZ-Gyro and Devico DeviGyro were employed for RC drilling during 2025. The downhole quality control checks were done at the drill rig, which included the survey reading and inspected the magnetic intensity, gravity, deviation and standard deviation of the survey performed with magnetic surrey tools. If all readings were within expected tolerance limits, the survey was accepted.

Downhole survey measurements were generally completed with the first reading at 12.5 m, followed by measurements at approximately 25 m intervals for the majority of drilling campaigns. More recent gyro surveys also include additional closely spaced survey stations, typically at 1 m intervals.

No documented procedures relating to magnetic corrections, magnetic susceptibility testing or independent verification surveys were provided. However, the progressive adoption of gyro survey instruments, particularly

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those referenced to true north, substantially reduces the potential influence of magnetic interference on recorded drillhole trajectories. Furthermore, SRK did not identify any material inconsistencies in drillhole trajectories during the three-dimensional (3D) validation of the drilling database.

Table 10‑1    Table showing drilling contractor and downhole survey used between 2022 and 2026 (as provided by source Aya)

Drilling Contractor

Year

Drill Type

Survey Tool

Manufacturer

Reference North

Geosond

2022

DD

EZ-Track

Reflex

Magnetic

Geosond

2023

DD

EZ-Track

Reflex

Magnetic

Geosond

2023

DD

DeviFlex

Devico

No

Geosond

2024

DD

DeviFlex

Devico

No

Geosond

2024

DD

DeviGyro

Devico

No

Geosond

2025

DD

DeviGyro

Devico

No

Geosond

2025

DD

GyroMaster

SPT

True North

Geosond

2,026

DD

GyroMaster

SPT

True North

FTE

2024

DD

EZ-Track

Reflex

Magnetic

FTE

2025

DD

EZ-Track

Reflex

Magnetic

FTE

2025

RC

EZ-Gyro

Reflex

True North

FTE

2025

RC

DeviGyro

Devico

No

FTE

2026

DD

EZ-Track

Reflex

Magnetic

10.5Hole Orientation

Drilling has generally been designed to intersect the mineralization as close to perpendicular to strike as practicable, with drillholes typically collared from the hangingwall where topography and site access permit. However, south of 3475177 N (in the south zone), the majority of the drillholes have been drilled from the footwall side of the mineralized system, as the steep and rugged terrain locally constrained drill pad locations and prevented drilling from the preferred orientation, resulting in locally sub-optimal intersection angles for certain drillholes. However, given the steeply dipping nature of the mineralization, and variation in dip most drillholes still intersect the mineralized zones at reasonable angles. Consequently, SRK considers that these locally sub-optimal intersection angles are unlikely to have a material impact on the geological interpretation or grade estimation presented in this report.

SRK recommends that future drilling programs continue to consider the orientation of the interpreted mineralized wireframes during Aya drillhole planning to maximize the angle of intersection and minimize excessively long, non-perpendicular mineralized intercepts where terrain and site access allows.

Figure 10‑2 presents a plan view of the drillhole traces relative to the interpreted mineralized wireframes, while Figure 10‑3, presents a representative cross-section illustrating drilling completed from both the hangingwall and footwall sides of the mineralized system.

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Figure 10‑2    Plan view showing the modelled wireframes (green) vs the drillhole traces at Boumadine (includes excluded drillholes)

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Figure 10‑3    Cross-sections A and B showing interpreted mineralized zones and drillhole traces. Section A is located at X = 317377, Y = 3476536; Section B is located at X = 317754, Y = 3473655. Drillhole traces are shown in pink, interpreted zones in green, and topography in grey.

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10.6Core Orientation

Core orientation was routinely completed during the modern diamond drilling programs to enable the collection of oriented structural data and improve the interpretation of the geological framework and mineralized structures. Review of the Project database indicates that the Company utilized both Devico DeviCore BTT and Imdexs ACTx (Figure 10‑4) core orientation systems during the exploration programs; however, no information was available to identify which drillholes were oriented using each device. Orientation marks were transferred onto the drill core immediately following retrieval from the drill string and maintained throughout the logging process to facilitate the measurement of geological structures.

This orientation line provided a reference which runs along the length of the core, relative to which the orientations of planar structural features (inclusive of bedding, foliation, veining, faulting, etc.) and/or linear features (lineations, striations, etc) could be determined.

For the orientated drilling, structural features such as faults, fractures and veins were recorded. The depths of the features were measured along with orientation data such as the alpha and beta angles on the core. The alpha angle is the acute angle described by the plane of the structural feature (the line joining the top and bottom closures of the core ellipsoid), and a line parallel to the core axis. The beta angle is the angle between the reference orientation line along the core and the bottom of the ellipse in a clockwise direction on the down dip end of the structural feature (0-360º).

Oriented structural logging largely depends on accurate marks from the orientation tool for downhole orientation. At Boumadine each structural measurement is assigned an orientation confidence rating ranging from 0 (low) to 4 (high), allowing lower-confidence measurements to be readily identified during subsequent structural interpretation.

Figure 10‑4    Core orientation procedure using the ACTx tool. A: ACTx orientation tool fitted to the core barrel with green status lights indicating successful operation. B: Bottom-of-hole orientation mark recorded on the end of the core run after ACTx orientation.

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10.7SRK Comments

SRK considers that the exploration work completed by Maya and Aya is appropriate for the style of mineralization and the current stage of the Project and conforms to industry best practices. The combination of geological mapping, hyperspectral interpretation, airborne geophysics, surface sampling and extensive drilling including orientated drilling has provided a suitable basis for geological interpretation and grade estimation. SRK notes that drilling remains the principal source of data supporting the 2026 MRE. Confidence in the geological interpretation decreases in areas of reduced sample coverage and is reflected in the classification of mineral resources, as noted in Section 14.12.

The procedure used to acquire drilling information in programs prior to Maya/Aya are not fully documented. Historical holes (surface and underground), make up 11% of the DD database. These are generally surrounded by newer Maya and Aya drillholes which largely confirm the location of mineralization which indirectly suggests that the location accuracy of the historical holes is reasonable. However, as noted in Section 14.1, the historical drillholes have been excluded from the 2026 MRE due to a lack of QAQC and or core available for verification sampling and assaying.

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11.0SAMPLE PREPARATION, ANALYSIS AND SECURITY

11.1Introduction

The following section relates to the logging and sampling methods and protocols used by Aya and Maya during the exploration conducted post 2017. Logging and sampling procedures have evolved during successive exploration campaigns; however, the overall workflow has remained consistent throughout the modern exploration programs.

The historical BRPM drilling programs pre-date the current exploration undertaken by Maya and Aya. Although historical geological and assay data have been retained within the Project database, very little documentation describing the associated logging, sampling, sample preparation, analytical and sample security procedures relating to the historical BRPM drilling has been made available to SRK. Consequently, this section primarily describes the procedures adopted during the modern Maya and Aya exploration programs.

For the DD, the drill core from each run was placed in the core boxes (plastic and wood) at the rig by the drilling team, with plastic spacers inserted at the start/end of each run, marked with the appropriate depth. At the end of each shift the drillcore is transported to Aya’ s secure core logging and sampling facility where the core is then subjected to the following steps:

•the core was reconstructed,

•geotechnically logged,

•Structurally logged (orientated core)

•geologically logged,

•photographed

•marked for sampling prior to being cut

•Core cutting and preparation of samples for dispatch.

For the RC, the drill cuttings were collected from the cyclone at nominal 1 m intervals. Further sample reduction was obtained using a riffle splitter, with one split retained for laboratory analysis and a second split retained in labelled chip trays for geological logging, photography and future reference.

11.2Core Recovery

For the Maya and Aya drilling core recovery was measured for each drilling run at Aya’s core logging and sampling facility after the core was reconstructed within the core trays using depth blocks to identify individual drill runs. The recovered core length was measured and compared with the drilled interval to determine the percentage core recovery. Core loss and Rock Quality Designation (RQD) were recorded concurrently within the Project database as part of the geotechnical logging process. No documentation relating to core recovery or RQD measurements for the historical BRPM drilling programs was available for review by SRK.

Review of the Project database indicates that core recovery information is available for 879 modern diamond drillholes, comprising approximately 115,900 individual core runs. The recorded dataset represents approximately 346,133 m of drilled core, with a very good overall average core recovery of 98.3%. A small proportion of individual core runs record recoveries greater than 100%, which are interpreted to reflect minor measurement tolerances, reconstruction of broken core and local core expansion following drilling, and are not considered unusual for diamond drilling programs.

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The drill core observed in the core boxes during the SRK site visit appeared to exhibit generally very high core recovery within both the fresh mineralized and unmineralized zones. Slightly lower, although still good, recoveries were observed within weathered and faulted zones, which are consistent with the core recovery database. SRK also reviewed a representative selection of drill core photographs as part of this study, which supports this conclusion. A representative example of the core recovery is presented in Figure 11-2.

No information relating to the total sample recovery for the RC drillholes has been provided to SRK, other than that a split of the drilled samples is weighed (typically 4.7 kg). SRK recommends that, for any future RC drilling program, the total drilled interval sample weight is recorded to determine a theoretical sample recovery based on the logged lithology.

Average core recovery remained consistently high throughout the modern exploration programs, varying between approximately 97.2% and 98.9% between 2022 and 2026. Local reductions in core recovery occur within zones of intense faulting, weathering and fractured rock; however, SRK did not identify any systematic relationship between reduced core recovery and mineralization that would materially affect the geological interpretation or grade estimation.

Figure 11‑1    Histogram showing the core recovery data provided to SRK for review

Figure 11‑2    Image of BOU-DD26-817, taken during the SRK site visit

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11.3Geotechnical Logging

Geotechnical logging was completed on all modern diamond drillholes prior to geological logging and sampling. Parameters recorded within the Project database include core recovery, Rock Quality Designation (RQD), fracture frequency, fracture characteristics and other observations relating to rock quality and ground conditions. These data support the geological interpretation and provide a dataset for geotechnical and mining studies.

No documentation is available describing the geotechnical logging procedures undertaken by BRPM on the historical drillcore, there are also no records associated with the 2017 Maya drilling.

Review of the Project database indicates that the average recorded RQD is 85.1% (median 93.2%), indicating generally competent rock conditions across much of the Project. Lower RQD values are typically associated with zones of faulting, weathering and locally fractured ground.

The geotechnical data collection to be appropriate for the current stage of exploration and suitable for geological interpretation.

11.4Geological Logging

At the Aya core shed, drill core was laid out in a series of core boxes on logging tables (Figure 11‑3). The core logging is qualitative in nature whereas the sampling and results this gives is quantitative in nature.

Geological logging was completed on all modern DD following geotechnical logging and prior to sampling. Logging was undertaken by Company geologists using standardised digital logging templates and controlled geological coding schemes, with Aya/Maya geological supervisors reviewing each drillhole to ensure consistency between drilling campaigns and adherence to the Company's logging standards. The geological database was developed and maintained using GeoticLog and subsequently exported for geological modelling and grade estimation.

Historical geological logs are available for a proportion of the BRPM drillholes; however, no documentation describing the historical geological logging procedures, coding schemes, logging standards or quality control procedures was available for review by SRK. As part of its data verification program, Aya relogged the available drill core from the 2018 to 2022 drilling programs during 2023 and also re-sampled selected intervals of the available drill core for verification purposes, as described in Section 12.4.

Geological logging included the systematic recording of lithology, weathering, alteration, mineralization, veining, oxidation, structure and other geological observations considered relevant to the interpretation of the deposit. Lithological contacts, mineralized intervals and major geological features were identified during logging and subsequently used to define the sampling intervals.

Aya provided SRK with its legends and codes associated with the geological logging, for the purposes of the 2026 MRE. It should be noted that massive sulphide (MS) is distinguished by approximately >70% logged visible sulphide, whilst semi massive sulphide is <70% and >30% logged visible sulphide. Oxidised mineralization (gossan) is denoted by A3 code.

During geological logging, the oriented core was used to collect structural data, such as the orientation of veins, faults, fractures, lithological contacts, foliations and other structural features (azimuth and dip). Each structural measurement is assigned an orientation confidence rating (0 to 4). This confidence ranking enables lower-quality orientation measurements to be identified and, where appropriate, excluded from structural interpretation and analysis.

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Geological observations and structural measurements formed an important component of the interpretation of the volcanic stratigraphy, synvolcanic fault architecture and controls on mineralization, and are subsequently incorporated into the geological interpretation supporting the 2026 MRE.

In addition to diamond drilling, geological logging was also completed on RC drillholes. Drill cuttings were collected from the cyclone and representative material from each sample interval was washed and placed into labelled plastic chip trays, with each compartment representing a discrete interval, typically 1 m in length. Company geologists examined the washed drill chips and systematically recorded lithology, weathering, oxidation, alteration, mineralization and other geological observations using the same controlled geological coding scheme adopted for the DD program. The chip trays were retained at the Project site as a permanent geological record and were available for subsequent review and validation.

During the site visit, SRK observed the DD geological logging procedures undertaken by the Company's geological staff and reviewed representative drill core and RC chip trays from a selection of modern drilling. SRK considers the geological logging procedures, geological coding schemes and quality of the geological database to be appropriate for the style of mineralization and sufficient to support the geological interpretation and grade estimation presented in this report.

Figure 11‑3    Aya core logging facility

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11.5Photography

DD core was systematically photographed following geological and geotechnical logging and after sample intervals had been marked, but prior to cutting and sampling. During the site visit, the Aya geological team advised SRK that each section of core is photographed twice: initially in its natural wet condition immediately following logging and subsequently after the core has dried. This procedure provides a permanent visual record of both the wet and dry appearance of the drill core and supports subsequent geological review and interpretation.

Drill core photographs were acquired using the Imago Capture software. Images were captured with up to three core boxes positioned on the photography table. Immediately following image acquisition, the photographs are reviewed within the software to ensure image quality and completeness (Figure 11‑4). Where issues such as poor focus, lighting or image quality are identified, the photographs are retaken before the core proceeds to the cutting and sampling stage. Following quality verification, the photographs are processed within the Imago software.

At the time of SRK's review, the Imago photographic database contained drill core imagery for the 2024, 2025 and 2026 drilling programs. Aya has informed SRK that they hold core photographs taken in 2022 and 2023 and plan to upload these to Imago in the coming months.

During the SRK QP site visit, SRK reviewed the Company's drill core photography procedures and subsequently examined representative drill core photographs during the desktop phase of the study. SRK also used the core photographs during the modelling of the weathering domains. Currently, RC chip trays are not photographed by the Company, and no core photographs exist prior to 2022. SRK recommends that Aya photograph all available drill core and RC chip trays and incorporate these records into the Imago database to establish a comprehensive digital photographic archive for the Project.

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Figure 11‑4    Aya core photography stand, with inlay image of drillcore image in Imago Capture software

11.6Density and Moisture Measurements

Density measurements were routinely completed on modern diamond drill core following the completion of geological and geotechnical logging and prior to sampling. Density determinations were undertaken using the Archimedes water displacement method, with the resulting measurements recorded within the Project. This procedure has been routinely undertaken since 2018.

During the SRK site visit, the Aya geological team demonstrated the density determination procedures. Density measurements were undertaken in a dedicated enclosed room to minimize environmental influences on the weighing process (Figure 11‑5). Prior to commencing daily measurements, the electronic balance was calibrated using certified 200 g and 500 g reference weights to verify the accuracy of the instrument. The water was also regularly changed according to Aya. The drillcore is first dried in an oven at 105°C, the dry core (mass) and submerged core (mass) is then measured to calculate the bulk dry density. SRK notes that the drill core was not wrapped in plastic film or coated with paraffin wax prior to measurement. In the opinion of the

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SRK QP, the majority of the drill core appeared sufficiently competent that such treatment (wrapping of drillcore) does not appear to be warranted. However, for locally broken or highly fractured core, SRK recommends that Aya consider the use of plastic film or paraffin wax coating to minimize the potential effects of water ingress during density determination.

The frequency of density measurements varied according to the lithology encountered. Within waste rock, density measurements were generally completed at approximately 10 m intervals. As massive sulphide mineralization was approached, additional density measurements were completed immediately prior to entering the mineralized zone, with measurements subsequently undertaken at approximately 1 m intervals throughout the massive sulphide intervals. However, density measurements within disseminated or apparently weakly mineralized zones were completed at a lower frequency. SRK recommends that the frequency of density measurements within these disseminated mineralized intervals be reviewed to ensure they are adequately represented within the density dataset. This may require additional density measurements to be collected following receipt of the assay results or, where appropriate, by re-sampling archived drill core. Density measurements were recorded within the Project.

SRK notes that the Company has not undertaken routine moisture content determinations on drill core samples. Although moisture content data are not required for the 2026 MRE, they represent an important physical property for future mining and processing studies, particularly the forthcoming Feasibility Study (FS). SRK recommends that Aya undertake a representative moisture content sampling program using fresh drill core collected from across the Project and analyzed within one to two days of drilling to minimize the effects of moisture loss. The sampling program should include representative samples of the principal lithologies, weathering states and mineralization styles expected to be encountered during future mining operations.

Figure 11‑5    Aya density measurement station

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11.7Sampling

Following geological logging, drill core was marked for sampling by Company geologists. Sample intervals were selected to honour geological contacts, lithological boundaries and mineralization, with a nominal sample length of approximately 1.0 m. Sample lengths generally ranged between 0.5 m and 1.2 m depending on the geological characteristics encountered.

During the SRK site visit, the Company advised that sample intervals are generated within dedicated sampling spreadsheets external to the geological logging database. Each sample is assigned a unique sample identification number, together with a corresponding Laboratory Information Management System (LIMS) code and QR code to maintain sample traceability throughout the analytical process, the sample information is entered directly into the GeoticLog database.

During the SRK site visit, the QP observed that, following core cutting, samples were placed into heavy-duty plastic sample bags. In addition to the sample ticket and QR code, the Sample ID was also handwritten on the exterior of each sample bag, providing a secondary means of sample identification should the QR label become damaged, detached or unreadable (Figure 11‑6 and Figure 11‑7).

RC samples were collected at nominal 1 m intervals directly from the cyclone. The cyclone discharged into a splitter, which produced three sample streams simultaneously: (i) a representative sample submitted for laboratory analysis, (ii) a duplicate sample collected at a nominal frequency of one in every 25 routine samples, and (iii) the remaining bulk reject, which was collected in large bags and retained at the drill site. All routine samples were weighed at the drill rig by an Aya technician using a calibrated hanging scale prior to dispatch (average weight of 4.7 kg). Representative samples were subsequently submitted for geological logging, sample preparation and laboratory analysis in accordance with the Company's sampling procedures.

No detailed documentation describing the sampling procedures adopted during the historical BRPM drilling programs was available for review by SRK. Consequently, the sampling methodology described below relates to the modern Maya and Aya exploration programs.

Figure 11‑6    Active sampling of drillcore at Aya core processing facility

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Figure 11‑7    Sampled drillcore with the sample identification and individual QR code denoted on the bag

11.8Core Cutting and Drying

Following sample interval selection, diamond drill core was cut longitudinally using a diamond saw (Figure 11‑8). One half of the core was retained within the core boxes (Figure 11‑6) as a permanent geological record, whilst the remaining half was submitted for laboratory analysis Quarter-core samples were routinely collected from selected intervals for insertion into the Company's Quality Assurance and Quality Control (QA/QC) program as field duplicate samples.

Following cutting, the samples were placed into individually labelled plastic sample bags and transferred to the on-site Afrilab sample preparation facility, the samples typically weighed between 2-4 kg. Prior to sample preparation, samples (including RC samples) were dried in accordance with the laboratory's standard operating procedures to remove moisture.

No documentation describing the core cutting, sample selection or sample drying procedures adopted during the historical BRPM drilling programs was available for review by SRK.

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Figure 11‑8    Aya core saw facility

11.9Sample Preparation

Since 2024, sample preparation has been undertaken at Afrilab's dedicated sample preparation laboratory located at the Project site. Prior to 2024, sample preparation was undertaken at Afrilab's laboratory in Marrakech. SRK has been provided with and reviewed Afrilab's documented sample preparation protocols and procedures.

No documentation describing the sample preparation procedures applied during historical BRPM drilling programs was available for review by SRK.

During the SRK site visit, the sample preparation procedures were demonstrated by laboratory personnel. Samples were initially crushed using a jaw crusher to achieve a target particle size of approximately 85% passing 2 mm. Following crushing, approximately 200 g of material was automatically split for further preparation. The remaining coarse reject was bagged, labelled and retained on site for future reference.

To minimize the potential for cross-contamination, compressed air was used to clean the jaw crusher between samples. In addition, clean sandstone was periodically passed through the crusher, followed by a further compressed-air clean before processing the next sample.

The split sample was subsequently pulverized using a puck pulverizer to achieve a target grind size of approximately 85% passing 75 µm (Figure 11‑9). Routine samples were pulverized for approximately four minutes, whereas blank (sandstone) material was pulverized for approximately one minute. Between each

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sample and blank, the pulverizer was cleaned using compressed air and a brush to minimize the potential for sample carry-over.

During the SRK site visit, laboratory personnel advised that sample preparation quality is routinely monitored through sieve testing (crushing and pulverizing).

Following pulverization, each pulp sample was visually inspected before being packaged into individual clean plastic bag envelopes. Approximately five analytical batches (~25 samples including QA/QC samples comprise each batch) were subsequently combined into a larger shipment bag before being transported to the Afrilab analytical laboratory in Marrakech for analysis, accompanied by a sample dispatch sheet.

Figure 11‑9    Rocklabs sample crusher

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Figure 11‑10    Pulverized samples being prepared for shipment to Afrilab analytical laboratory in Marrakech

11.10Sample Analysis

For the Maya and Aya drilling, the Company contracted Afrilab analytical laboratory in Marrakech to conduct the chemical analysis. No documentation describing the analytical laboratories, analytical methods, digestion techniques, analytical finish, detection limits or quality control procedures used during the historical BRPM drilling programs was available for review by SRK.

The analytical certificates reviewed by SRK indicate that Afrilab is an ISO 9001-certified laboratory and holds Canadian Association for Laboratory Accreditation (CALA) accreditation.

Au was analyzed by 50 g fire assay with an atomic absorption spectrometry (AAS) finish (Afrilab method MO/ANA/001). Ag, Cu, Fe, Pb and Zn were analyzed following aqua regia digestion (Afrilab method MO/ANA/002) with an ICP-OES finish. Samples returning Cu, Fe, Pb or Zn grades greater than 1% were routinely re-analyzed by atomic absorption spectrometry (AAS) to provide accurate assays at higher concentrations. Ag was initially analyzed by ICP-OES; however, since 2022 samples returning Ag grades greater than or equal to 200 g/t were routinely re-analyzed by 50 g fire assay with a gravimetric finish to provide accurate assays for high-grade mineralization. Aya confirmed that the fire assay gravimetric method has an upper analytical limit of 100% Ag (pure silver) and 100 g/t for Au and lower detection limit of 0.05 g/t for Au.

Additional trace elements, including As, Mo, Sn and S were analyzed following aqua regia digestion with ICP analysis, though the amount of assay associated with these is variable. For instance only 92 S% assay are contained within the database compared to 163,822 As values and 329,370 Mo values. The analytical methods provided lower detection limits of 1 g/t Ag, 0.01% Cu, Fe, Pb and Zn. While the routine ICP-OES method has an upper analytical limit of 1% for Cu, Pb and Zn, above which samples are routinely re-analyzed by AAS (no upper detection limit provided by Aya).

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Upon review of the Afrilab analytical certificates, SRK noted that base metal analyses have been undertaken using an aqua regia digestion method since 2017, rather than the four-acid digestion method previously reported in Boumadine technical reports.

In addition, an independent umpire laboratories has been used, during the 2022 to 2024 drilling campaigns. Umpire analyses were completed by ALS Seville, Spain, while from late 2024 onwards this work transitioned to ALS Morocco (Zgounder) following commissioning of this laboratory. In addition, independent verification samples were submitted to Actlabs (Ancaster, Canada) by P&E during its 2024 due diligence program as part of an external verification exercise.

11.11Sample Security

Throughout the exploration programs, drill core remained under the control of Company and drilling contractor personnel from the drill rig to Ayas secure core logging facility. Following logging, photography and sampling, prepared samples were securely transported to the on-site Afrilab sample preparation laboratory. Following sample preparation, analytical pulps were packaged into sealed shipment bags and transported to the Afrilab analytical laboratory in Marrakech for analysis. Coarse reject material was retained, bagged and stored at the Project site for future reference.

No documentation describing the chain of custody or sample security procedures employed during the historical BRPM drilling programs was available for review by SRK.

The use of unique sample identification numbers, QR codes and the Company's Laboratory Information Management System (LIMS) provides traceability throughout the sampling and analytical process.

11.12SRK Comments

SRK reviewed the Company's geological logging, sampling, sample preparation, sample security and analytical procedures during the site visit and considers these to be appropriate for the style of mineralization and that they are generally aligned with industry-standard practices, and adequate for the purposes of the 2026 MRE. The dedicated on-site sample preparation facility and routine equipment cleaning procedures provide additional confidence in the quality and integrity of the prepared samples submitted for laboratory analysis.

The procedures used for logging, sampling, sample preparation, analytical testing and sample security during historical BRPM drilling programs are not adequately documented. In addition, limited information is available to SRK regarding the drilling, logging, sampling, sample preparation and QA/QC procedures employed during the 2017 Maya drilling program. Furthermore, the drill core from the 14 Maya drillholes completed during 2017 could not be located within Aya's core storage facilities and was therefore unavailable for verification by relogging or re-sampling.

Although historical and 2017 drilling data is available within the Project database, the absence of documented QA/QC procedures, together with the lack of archived drill core and supporting records available for independent verification, limits confidence in the historical analytical data. Historical and 2017 drillholes are generally surrounded by modern Maya and Aya drillholes, which largely confirm the geometry and tenor of the mineralization, though there is some uncertainty over the spatial position of the historical drilling as this data has been translated into a UTM system from the historical mine grid.

SRK recommends that Aya continue to maintain and periodically review its sampling and QA/QC procedures as exploration progresses. In addition, SRK recommends that representative moisture content measurements be undertaken on fresh drill core during future drilling programs to support the forthcoming FS. Samples

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selected for moisture determination should be representative of the principal lithologies, weathering states and mineralization styles expected to be encountered during mining operations.

12.0DATA VERIFICATION

12.1Introduction

SRK undertook a program of independent data verification to assess the suitability of the geological, drilling, sampling, analytical and survey data supporting the MRE. The verification program included a site visit; review of the drilling and assay databases; assessment of the QA/QC data and Aya's drill core re-sampling program; verification of selected drillhole collar locations; assessment of historical underground mine plans and depletion surveys; review of drill core orientation procedures; and comparison of the geological interpretations against the available drilling, surface and underground geological information.

The verification program was designed to provide confidence that the Project database is suitable for use in a MRE. Particular attention was given to the quality and reliability of the modern Maya and Aya drilling database, which forms the basis of the 2026 MRE. Historical BRPM drilling and underground information were also reviewed where appropriate.

In addition, SRK has reviewed the findings and recommendations presented within previous technical reports where relevant to the current commission but it has not relied solely upon the conclusions of previous consultants in preparing the 2026 MRE.

12.2SRK Verification

SRK completed both desktop and field-based verification of the geological, drilling, survey and analytical information supplied by the Company. Verification activities included review of the Project database, inspection of drill core, observation of logging and sampling procedures (including drillhole logs), verification of selected drillhole collar locations, comparison of historical underground workings with current topography and review of the core orientation methodology.

The onsite verification work was completed by Mr Guy Dishaw, P.Geo., QP, during the site visit, together with subsequent desktop reviews undertaken by the SRK project team.

12.2.1Site visit

Guy Dishaw (QP) completed a site visit to the Boumadine Project between 23 and 28 April 2026. The purpose of the visit was to verify the geological setting, observe current exploration practices and assess the quality of the geological, drilling, sampling and analytical information supporting the Mineral Resource Estimate (MRE).

During the site visit, the Qualified Person:

•observed active diamond drilling operations and reviewed drilling practices;

•reviewed the drill core handling, reconstruction, orientation, geotechnical logging, geological logging and sampling procedures;

•observed the density determination and sample preparation procedures (this was limited);

•inspected representative diamond drill core and RC chip trays;

•independently verified the locations of 14 selected drillhole collar locations (Figure 12‑1);

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•undertook high-level confirmatory geological mapping of representative surface exposures, including gossanous outcrops and historical artisanal mining excavations, to confirm the geological interpretation and styles of mineralization and orientation of structures (Figure 12‑2).

•compared the geological interpretation (Aya geological logs) against drill core observations and surface exposures;

•surveyed the locations of selected historical mining infrastructure, including shafts and adits, using a handheld Global Positioning System (GPS) receiver to assess the accuracy of the georeferenced historical mine plans;

•reviewed the core photography procedures and the digital photographic archive;

•inspected the on-site Afrilab sample preparation facility at Boumadine and observed the sample preparation workflow;

•undertook a high-level inspection of the Afrilab assay laboratory in Marrakech and observed the laboratory workflow.

•discussed the geological interpretation, exploration programs and data management procedures with Company personnel.

During the collar verification component of the SRK site visit, only one of the 14 selected historical drillhole collars could not be directly located due to the absence of an identifiable collar marker. However, field evidence, including the presence of a drill sump and broken concrete, provided some confidence in the recorded collar location. The remaining selected drillhole collars were located and verified in the field by SRK using a handheld GPS. Based on the results of the collar verification program, SRK considers the drillhole collar locations to be sufficiently accurate for the purposes of geological modelling and the 2026 MRE.

The high-level geological mapping undertaken during the SRK site visit provided an independent assessment of the geological framework. The observed lithologies, alteration assemblages, gossan development, surface structural features, including those exposed within artisanal workings, and styles of mineralization were consistent with the geological interpretation adopted for the 2026 MRE and with the geological logs selected for review. Structural observations made at surface were also consistent with structural measurements obtained from oriented drill core and with the interpreted orientations of the principal mineralized structures.

Overall, SRK considers that the geological and structural observations made during the site visit are consistent with the information contained within the Project database.

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Figure 12‑1    Drillhole BOU-DD95-404 verified by SRK in the field

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Figure 12‑2    Gossanous outcrop observed at surface during the high level geological mapping undertaken by SRK

12.2.2Twin drillhole Comparison

No twinned drillholes have been completed at Boumadine, however a few drillholes were terminated due extreme deviation and a later hole was redrilled covering a similar area, however in all circumstances the terminated drillholes never intersection modelled mineralization (included in the 2026 MRE wireframes). There are also a few sections of drillholes which occur within close proximity to one another (drillholes drilled at different collar sites with different azimuth and dips).

SRK recommends that twinned drillholes be completed at several representative locations across the deposit. Where practicable, the program should include twins of each drilling method used at the Project (MP, RC and DD) by DD. The twin DD drillholes should be used to assess the precision and representativeness of the different drilling methods through visual and statistical comparisons. The additional closely spaced sample pairs would also assist in quantifying short-scale grade variability (nugget effect) and support geostatistical analysis (improve confidence in the interpretation of short-range grade continuity).

12.2.3Verification of underground depletion as-builts

The historical underground workings were captured in a local Mine grid; the supporting mine plans and associated as-builts were transformed from the historical local mine grid into the Project UTM coordinate system by GoldMinds Geoservices Inc.

SRK reviewed a selection of the georeferenced historical mine plans used by GoldMinds in the construction of the depletion meshes and considers that these provide an appropriate representation of the overall historical mining geometry for geological interpretation and depletion purposes. However, numerous local positional discrepancies remain evident between some of the georeferenced underground workings and the mineralization wireframes interpreted as part of the 2026 MRE.

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As part of the verification process, SRK compared the locations of selected historical shafts, adits and other surface mining infrastructure shown on the georeferenced mine plans with their observed positions in the field using a handheld GPS receiver. These observations were undertaken in conjunction with the geological mapping program and provided an independent assessment of the positional accuracy of the georeferenced historical mine plans and associated depletion surveys. The review indicated that the overall georeferencing is suitable for the purposes of geological interpretation and depletion, with the locations of the selected surface features generally within the expected positional accuracy of the handheld GPS receiver.

However, local positional discrepancies remain between some of the georeferenced historical mine plans and as-builts in relation to the modern drilling database used to interpret the mineralization wireframes. Consequently, the historical underground information, including the interpreted orientations and extents of the stopes, can only be used as a guide for geological modelling, with greater reliance placed on the Aya drilling database where inconsistencies between the datasets were identified. Additional depletion adjustments were therefore required to account for mineralization interpreted to have been historically mined but which does not spatially coincide with the available depletion meshes, as discussed in Section 14.13.

12.2.4Inspection of Afrilabs Analytical laboratory

As part of the data verification program, SRK also completed a high-level inspection of the Afrilab sample analytical laboratory in Marrakech on the 28 April 2026. The purpose of the visit was to gain an understanding of the laboratory layout, sample handling procedures and analytical facilities used for the Project. During the visit, SRK observed that the laboratory was clean, well-organised and logically laid out, with samples stored in a neat and clearly labelled manner throughout the analytical workflow. The QP viewed the fire assay furnaces and ICP analytical machines (Figure 12‑3) and discussed the laboratory procedures, analytical methods and quality control protocols with the laboratory manager. Although the inspection did not constitute a formal laboratory audit, no issues were identified that would materially affect confidence in the analytical results reported for the Project.

Figure 12‑3    Fire assay furnaces (A) and the inductively coupled plasma (ICP) analytical instrument (B) at the Afrilab laboratory, Marrakech, Morocco.

12.2.5Verification of sample database

SRK completed a phase of data validation on the digital sample data supplied by the Company, which included the following:

•search for sample overlaps, duplicate or absent samples;

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•checks for anomalous assay results;

•checks for incorrect or irregular survey results;

•search for non-sampled drillhole intervals within the mineralized zones; and

•Review of the exported database and cross-check against a selection of the 2025 Afrilabs assay certificates, comprising 485 Au samples, 778 Ag samples, and 992 Cu, Fe, Pb, and Zn samples.

SRK did not identify any material issues during the data verification process. However, several minor database issues were identified, including missing drillhole type codes and inconsistent formatting of analytical detection limits (e.g. spaces between the "<" and ">" symbols and reported values), as well as 6 samples reporting Ag at 200 g/t using ICP, these all relate to tailings drilling which is does not form part of this MRE. These issues were communicated to Aya and were corrected by Aya.

12.3Verification by the Company and External Consultants

In addition to the independent verification completed by SRK, the Company has implemented a series of internal verification procedures throughout the exploration programs to ensure the quality and integrity of the Project database. These procedures include geological supervision (Aya personnel) of drillhole logging, routine validation of geological and sampling information prior to incorporation into the Project database, routine review of QA/QC performance, and ongoing reconciliation of geological interpretations against newly acquired drilling data.

As part of the redevelopment of the Boumadine Project, Maya commissioned GOULEX, a French exploration consulting company, in 2014 to compile and validate the available historical exploration and mining data. The scope of work included the verification and surveying of historical drillhole collars, shafts and tailings facilities using differential GPS (DGPS), with all surveyed coordinates recorded in the Project UTM coordinate system. The historical underground mine plans and associated underground drilling data were subsequently transformed from the historical local mine grid to the Project UTM coordinate system by GoldMinds Geoservices Inc., providing the basis for incorporation of the historical information into the modern geological database.

The Company has also commissioned a number of independent technical reviews of the Project during successive exploration programs. Most notably, P&E Mining Consultants Inc. completed independent reviews of the geological database, sampling procedures, analytical QA/QC program and MRE, as documented in the 2024 NI 43-101 Technical Report. The P&E study also included an independent QP site visit, as well as the collection and submission of independent verification samples for laboratory analysis at Actlabs in Ancaster, Canada.

The Boumadine Project has also been the subject of additional independent technical studies, including reviews of the geological setting, geological mapping and other supporting technical information relevant to the 2026 MRE.

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12.4Aya Drillcore re-Sampling

All available drill core from the 2018 to 2021 drilling programs was relogged by Aya during 2023, and selected intervals were resampled during the same time period. The resampled intervals were derived from quarter-core opposed to the half-core samples (original sample). Drill core from the earlier programs (2017 Maya and BRPM) could not be verified through resampling or relogging, as the core could not be located. The resampling program was undertaken due to the lack of accompanying QA/QC samples associated with the original assays, both in terms of the types of QA/QC samples inserted and their frequency of insertion. The resampled core was submitted for analysis together with accompanying QA/QC samples (CRM’s, duplicates, blanks), see Section 12.5. Aya informed SRK that, where reassaying and relogging were completed for the post-2018 drillholes, the reassayed and relogging data were used in place of the original assay and logging data within the Project database. Aya also informed SRK that in some cases the resampling intervals did not match the original sample intervals.

SRK was provided with Aya's comparative statistics Excel spreadsheet, which compares the original assay results with the corresponding results from the resampled drill core from 10 drillholes drilled between 2018-2020 (Maya drilling). It should be noted that SRK did not independently review the resampling intervals during the site visit and has not observed the resampling program. Figure 12‑4 to Figure 12‑7 are scatter plots for showing the resampling results for Au, Ag, Fe and Pb.

A total of 207 intervals contained paired original and reassayed Au results, while 170 samples contained paired Ag results, 173 contained paired Zn results, and 72 contained paired Fe and Pb results, other analytes were also analyzed but not compared by Aya graphically.

The samples were selected across a range of grades from both visually identified mineralization and from the surrounding lower grade and waste units. Aya compared the original and reassayed results using scatter plots to provide a visual assessment of the correlation between the two datasets. Aya and the previous consultant, P&E, concluded that the drill core resampling results were of sufficient quality, with minimal bias observed and that these could be used to support an MRE.

SRK review of Aya's comparative statistics spreadsheet and notes that the total number of paired samples available for verification is relatively low for a comprehensive assessment of the original assay data. However, the number of samples is considered sufficient to identify any material bias between the original and reassayed datasets. Overall, the scatter plots show greater variability for Au compared with Ag, Fe and Pb (Figure 12‑4 to Figure 12‑7). Several anomalous results were identified, including one reassayed sample that returned 61.44% Pb compared with an original assay result of 0.06% Pb (note this is not included in Figure 12‑7 due to scaling of Y axis). The same sample returned a reassayed Ag grade of 2,176 g/t compared with an original grade of 16 g/t, with similar differences also observed for the other analytes. The reason for these differences is unknown, although they may reflect differences between the original and resampled core interval and/or less likely the heterogeneous nature of the mineralization. SRK recommends that Aya investigate the reason for these discrepancies with the Ag assay values.

In addition, 19 samples returned reassayed Fe grades ranging from 3.33% to 25.38%, whereas the corresponding original samples all reported Fe values of 0.5%. The reason for these differences is unknown and again this may reflect differences between the original and resampled core interval and/or the heterogeneous nature of the mineralization. SRK recommends that Aya investigate the reason for these discrepancies with the Fe assays.

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Figure 12‑4    Scatter plot showing Aya resampled drillcore (Au g/t)

Figure 12‑5    Scatter plot showing Aya resampled drillcore (Ag g/t)

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Figure 12‑6    Scatter plot showing Aya resampled drillcore (Fe %)

Figure 12‑7    Scatter plot showing Aya resampled drillcore (Pb%)

Overall, the Pb data show reasonable correlation, although greater variability is observed at elevated Pb grades, typically above 1%. The Fe data show good correlation across the assayed grade range, excluding the 19 samples that returned reassayed Fe grades significantly higher than the corresponding original values. The Ag data also shows good correlation, with most samples returning grades below 100 g/t, excluding the anomalous sample discussed above. The Au data show greater scatter but remain within acceptable tolerances, with the grades encountered typically below 0.6 g/t. For the analytes reviewed by SRK, there

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appears to be a slight overall bias towards higher grades in the reassayed results. These biases are not considered material to the use of the data in the current study. Should further drill core resampling be undertaken, SRK recommends that the resampling intervals correspond as closely as practicable to the original sample intervals to improve the comparability of the datasets.

Overall, SRK considers the correlation between the original and reassayed datasets to be within acceptable tolerances, taking into consideration the different sample types (quarter core versus half core), potential different sample intervals, the relatively limited number of paired samples and the nature of the mineralization. SRK recommends that Aya investigate the reasons for the significant differences associated with the 19 Fe sample pairs discussed above, together with the anomalous Pb and Ag results. The biases observed between the datasets are not considered material to the 2026 MRE, and the overall findings are consistent with the conclusions previously reached by Aya and P&E (suitable for inclusion in a MRE).

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12.5Quality Assurance and Quality Control (QA/QC) Analysis

12.5.1Introduction

QA/QC procedures are implemented in modern exploration programs to assess the accuracy and precision of the sampling process and to check for bias and calibration of the sampling and analysis procedures. The accuracy is a measure of the sample bias and the precision, its repeatability. Poor accuracy can often be attributed to the instrument calibration, and poor analytical techniques. Whereas poor precision is often due to sampling or preparation procedures. The accuracy and precision of any sampling program can be monitored by a series of checks which may include, but are not limited to, the insertion of:

•Blank samples: Typically, these samples have a very low (known) grade of the elements of interest and are used to assess contamination between samples in laboratories.

•Certified Reference Material (CRM): are material where the grade(s) of the material has been determined by multiple laboratories by round-robin analysis and have an accompanying certificate.

•Duplicate samples including:

•Field duplicates – in core drilling, this typically represents a second sample from the same interval, typically these are quarter core samples collected from the retained half-core. These duplicates are used to test the effectiveness of the sampling method.

•Coarse duplicate – sample prepared from the reject coarse material created during sample preparation. These samples are used to check the homogeneity and effectiveness of the sample crushing and splitting stage.

•Pulp duplicate – sample prepared from the reject pulverized material created during sample preparation. These samples are used to check the homogeneity and effectiveness of the sample pulverization stage.

•Umpire Duplicate – re-analysis of the same sample (typically pulp) using a third-party laboratory. These are used to check for issues within the laboratory or analytical equipment which may introduce bias.

It should be noted that the blank and CRM datasets provided by Aya included analytical results for Au, Ag, Cu, Pb, Zn and a range of additional multi-element analyses. In contrast, the duplicate datasets supplied to SRK contained results for only Au, Ag and Mo. Consequently, SRK's assessment of duplicate sample precision was limited to these elements, other than Mo which has not been considered in SRK’s QAQC review. For completeness, SRK recommends that Aya include the full suite of analyzed elements, though as a minimum Cu, Pb and Zn should be included in the duplicate QA/QC database prior to the next MRE update. The insertion rates presented by SRK below are based on the number of Au samples associated with the QA/QC and assay database, which is biased by the resampling. Aya informed SRK that for a batch of 25 samples: one certified reference material (CRM), one blank and one field duplicate is typically inserted into the sample stream. Coarse duplicates and umpire samples are also analyzed, though they are fewer in number.

As noted above, all historical drill core from the 2018 to 2021 drilling programs was relogged and resampled with QA/QC samples by Aya during 2023, whilst the previous drilling (2017 Maya and historical BRPM) was not able to be verified through resampling. The 2018 Maya drilling program comprises the earliest drilling incorporated into the 2026 MRE and was either completed under, or subsequently resampled using, Aya modern sampling and QA/QC protocols. Consequently, these data are considered directly comparable with the subsequent Aya drilling programs and have been assessed as part of the current QA/QC review.

SRK did not review the internal laboratory QA/QC samples as part of this study. Instead, SRK's review focused on the Company's own QA/QC program.

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12.5.2Pre 2018 QA/QC

The limited drilling completed by Maya during 2017 predates the implementation of the current QA/QC program. The original drill core from the 2017 Maya program and the historical BRPM drilling programs could not be located during the current study, and the Company advised SRK that little to no routine QA/QC data are associated with the original analytical results. As the 2017 Maya and historical BRPM drillholes were excluded from the 2026 MRE, the associated analytical data have not been included in SRK's QA/QC assessment presented below. This approach is consistent with that adopted in previous studies.

12.5.3Blanks

Blank samples were routinely inserted into the sample stream throughout the modern exploration programs by Aya technicians and geologists prior to sample preparation, to monitor potential contamination during sample preparation and laboratory analysis. Blank samples were inserted at an approximate rate of 4.5%. In total, 17,206 blank samples are recorded in the QA/QC database for the period from 2018 to 2026.

The Company utilized two types of blank material. The primary blank material comprised local sandstone (labelled as Gris) and was included in all sample streams between 2018 and 2026. A second blank material comprised concrete, which was also periodically inserted into the sample stream between 2018 and 2022. Neither of the blank materials used by the Company has been independently certified.

SRK reviewed the analytical performance of the blank samples for Au, Ag, Cu, Pb and Zn. The majority of blank samples returned assay values at, or close to, the respective analytical lower detection limits, indicating that contamination during sample preparation and laboratory analysis is generally well controlled and there no evidence of systematic bias (Figure 12‑8).

Occasional isolated elevated values were identified; however, these occurrences were infrequent and did not indicate systematic contamination within the analytical program.

Although the analytical performance of the blank samples is considered satisfactory and raise no QA/QC concerns, SRK recommends that Aya consider introducing a commercially certified blank material with documented concentrations for the principal elements of interest, or alternatively undertake a round robin program on the locally sourced sandstone blank to certify this.

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Figure 12‑8    Blank performance between 2018 and 2026 (A= Au, B = Ag, C= Cu, D= Pb and E = Zn)

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12.5.4Certified reference materials

CRM’s sourced from OREAS were routinely inserted into the sample stream throughout the modern exploration programs (including the resampling program) to monitor the analytical accuracy and precision of the primary laboratory across a range of grades and material types.

Review of the CRM database indicates that Aya has utilized 31 different certified reference materials throughout the modern exploration programs. Collectively, the CRM suite provides reasonable grade coverage for the principal analytes of economic interest, (Au, Ag, Cu, Pb and Zn), encompassing both low-grade, medium-grade and high-grade mineralization expected to be encountered at the Project. Several of the CRMs were specifically developed for monitoring either precious metals (Au and Ag) or base metals and Ag and, consequently, do not contain certified values for all analytes or contain certified concentrations below the analytical reporting limits or above the grade encountered in the drilling database. Of the 31 CRMs reviewed by SRK, 11 are not certified for Au, one for Ag, and one for Zn.

Where certified values were unavailable, or where certified concentrations were below the analytical detection limits, SRK excluded those analytes from its assessment. Sufficient alternative CRMs containing appropriate certified concentrations are available to adequately assess laboratory performance across the expected grade ranges.

SRK assessed the analytical performance of the CRMs by comparing the reported assay values against the certified values and associated standard deviations supplied by OREAS for the appropriate digestion method. Results were considered acceptable where the reported assay value fell within ±2 standard deviations (SD) of the certified value, whilst results falling outside ±3 SD were considered failures requiring further investigation.

SRK completed an initial review of the CRM database supplied by Aya, which identified numerous instances where CRM sample IDs had been incorrectly assigned (mislabelled) or swapped within the database. The majority of these involved similarly named standards, such as OREAS 353 and OREAS 353B. These discrepancies were communicated to Aya during the review process and, at the time of preparing this report, the majority of the identified issues had been corrected and are therefore not presented in the figures below. A small number of sample swaps still remain in the database such as OREAS 234B and OREAS 236, these are either from the mislabelling of original samples or from samples being incorrectly assigned from other CRM’s or QA/QC sample types, SRK recommends Aya investigate these and attempt to resolve them if possible.

For Au, SRK used the certified 50 g fire assay values and associated standard deviations provided by OREAS. For Ag, SRK used the aqua regia digestion certified values and associated standard deviations for CRMs with certified Ag concentrations below 200 g/t, consistent with the routine analytical method employed by Afrilab. For CRMs with certified Ag concentrations greater than 200 g/t, SRK used the fire assay certified values and associated standard deviations, reflecting the ore-grade analytical procedure adopted by Aya. For Cu, Zn and Pb, SRK used the certified values and associated standard deviations for aqua regia digestion, where available. SRK notes that not all CRMs used by Aya are certified for aqua regia digestion for Ag <200 g/t or the base metals (Cu, Pb and Zn %), with some instead certified using other digestion methods, typically four-acid digestion. Where certified values for aqua regia digestion were not available but certified values for four-acid digestion were available, SRK assigned a lower level of confidence to the assessment of CRM performance, as the certified values are not directly comparable with the aqua regia digestion method employed by Afrilab. Nevertheless, no material issues were identified in the performance of these CRMs based on comparison with the four-acid certified values.

The CRMs that do not have certified values for fire assay analysis of Au and/or Ag (for Ag concentrations above 200 g/t), or for the base metals (Cu, Pb and Zn) using aqua regia digestion, include OREAS 100A (last inserted into the sample stream in June 2023), OREAS 354 and OREAS 994. SRK recommends that these CRMs are no longer inserted into the sample stream, as other CRMs certified for the applicable analytical methods adequately cover the typical grade ranges encountered at the Project. The exclusion of these CRMs is not considered material to the outcome of the QA/QC review. Furthermore, no material issues were

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identified in the Cu, Pb and Zn results for OREAS 354 and OREAS 994 when assessed against the available certified values, other than a small number of apparent sample swaps or mislabelling issues.

Overall, the CRM results demonstrate acceptable analytical accuracy for the principal elements of economic interest. The vast majority of analyses fall within the expected control limits, with only a limited number exceeding the ±3 SD failure limits. A number of these failures are interpreted to relate to historical CRM sample ID assignment errors. For example, the Au results for OREAS 234B suggest that several samples may have been incorrectly assigned and should instead correspond to OREAS 236. The certified base metal concentrations for this CRM are below the Afrilab reporting limits.

Several of the lower-grade CRMs exhibit comparatively poorer performance, with results falling outside the ±3 SD control limits. These CRMs generally contain analyte concentrations close to, or below, Afrilab lower detection limits, where analytical precision is expected to decrease. Although these CRMs locally exceed the QA/QC control limits, the absolute analytical differences are small and are not considered material to the assessment of laboratory performance or the 2026 MRE. Furthermore, analytical performance improves at higher analyte concentrations, providing confidence in the laboratory's performance across the grade ranges of principal economic interest.

For Pb, SRK notes that OREAS 353, is certified for a classical wet chemistry digest, and has returned a number of failures during 2022 and 2023. However, this CRM does not have a certified Pb value based on aqua regia digestion and is therefore not directly comparable with the aqua regia digestion method employed by Afrilab. Furthermore, the certified Pb concentration of 61.26% is approximately twice the maximum Pb grade contained within the Project assay database and is therefore not representative of the grades typically encountered during routine exploration or within the MRE dataset. Other Pb CRMs certified using aqua regia digestion, which cover grade ranges more representative of the Project mineralization, demonstrate acceptable analytical performance. In addition, OREAS 353 does not appear to have been routinely inserted into the sample stream after 2024.

SRK also identified two Zn CRM results reporting values greater than 100% (123% and 1299%). These values are considered anomalous and are likely to represent transcription or database errors. SRK recommends that Aya investigate and correct these records.

SRK has presented multi-CRM plots below (with rescaled y axis) in Figure 12‑9 to Figure 12‑13, which correspond to dates between 2022 to 2026, though all CRM plots are included in appendix A.

Overall, SRK considers the CRM program to demonstrate acceptable analytical accuracy and precision throughout the modern exploration programs. Following correction of the identified CRM sample ID discrepancies, no material issues were identified that would adversely affect confidence in the analytical database or the 2026 MRE.

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Figure 12‑9    Ag multi CRM plots (A = full range, B = rescaled Y axis)

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Figure 12‑10    Au multi CRM plots (A = full range, B = rescaled Y axis)

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Figure 12‑11    Cu multi CRM plots (A = full range, B = rescaled Y axis)

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Figure 12‑12    Pb multi CRM plots (A = full range, B = rescaled Y axis)

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Figure 12‑13    Zn multi CRM plots (A = full range with sample reporting 1299% Zn omitted, B = rescaled Y axis)

12.5.5Duplicates

Duplicate samples are intended to assess the repeatability of the analytical results and provide a measure of the precision associated with the sampling, sample preparation and analytical procedures. Comparison of the different duplicate sample types can also be used to identify potential issues within the sample preparation workflow, including crushing, splitting and pulverization.

SRK notes that Aya does not routinely include pulp duplicate samples within its QA/QC program, instead it typically includes field and coarse rejects within the sample stream. SRK recommends that pulp duplicates be incorporated into future QA/QC programs, as they provide an additional measure of the homogeneity of the pulverized sample and the effectiveness of the crushing and pulverization stages. Field duplicates typically exhibit greater variability than coarse reject or pulp duplicates because they incorporate the additional

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variability associated with sample collection. Consequently, pulp duplicates can provide a more sensitive assessment of laboratory precision and sample preparation performance.

SRK acknowledges that the Company's umpire laboratory program involves the re-analysis of prepared pulp samples and, therefore, provides some information on pulp homogeneity and analytical repeatability. However, the primary purpose of umpire samples is to assess analytical accuracy and identify potential bias between laboratories, rather than to routinely monitor the precision of the primary laboratory's sample preparation procedures.

SRK identified 38 duplicate sample pairs for which the duplicate sample has no reported Au assay (blank entry in the database), despite the corresponding original sample having a reported Au assay. This issue was not observed for the Ag duplicate dataset. SRK recommends that Aya investigate these records and or provide an explanation on this.

SRK also identified a number of Au and Ag duplicate pairs where assay values of zero associated with the field, coarse reject and umpire duplicate database. Specifically, 44 zero values occur within the duplicate assay column and 191 within the original assay column, despite the corresponding paired duplicate assay containing a value for the same element. Although these zero values are likely to represent assays below the analytical detection limit, this could not be confirmed from the duplicate database alone. SRK recommends that Aya review these records against the original analytical certificates and replace the zero values with the reported assay values or, where appropriate, the laboratory detection limit. Where no analytical result is available, the assay value should remain blank and the reason documented within the database, rather than assigning a value of zero.

As noted previously, SRK's assessment of the duplicate database was limited to Au and Ag, as duplicate assay results for the remaining analyzed elements (Cu, Zn and Pb) were not provided by the Company.

12.5.6Field duplicates

Field duplicate samples (recorded as COARSE_DUP_UMP by Aya) were routinely inserted into the sample stream to assess the precision of the complete sampling, sample preparation and analytical process. Overall, field duplicates were inserted at an average insertion rate of approximately 3.41% (13,019 total samples) and have date range between with 2023 and 2026. A small portion of these samples relate to the Aya resampling and the increase of additional QAQC samples.

For DD, field duplicates comprised quarter-core duplicate samples, whereby one-quarter of the drill core was submitted as the duplicate sample following routine half-core sampling. For reverse circulation (RC) drilling, field duplicates comprised second representative splits collected directly from a riffle splitter at the drill rig. These duplicate samples were intended to assess the precision associated with both the sampling procedures and the laboratory analytical process.

SRK reviewed the field duplicate results for Au and Ag via scatter graphs (Figure 12‑14) and considers in general the duplicates are in reasonable agreement with their original sample grades, as noted by the moderate to high statistical correlation for Ag (R2= 0.83) and Au (R2=0.80), and visual inspection of the graphs. The coarse reject duplicate graphs are also impacted by zero values in the original value, though this will not impact the R2value. The observed variability is consistent with that expected for polymetallic mineralization and is considered acceptable for the 2026 MRE.

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Figure 12‑14    Field duplicate performance between 2018 and 2026 (A = Au and B = Ag)

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12.5.7Coarse reject duplicates

Coarse reject duplicate samples are intended to assess the repeatability of the laboratory analyses while also providing a measure of the consistency of the crushing and sample splitting procedures. As the duplicate samples originate from the same sample, they can also be used to identify potential issues associated with sample preparation prior to analysis.

Coarse reject duplicate samples (recorded as DUP by Aya) were routinely included within the QA/QC program between 2018 and 2024 to assess the repeatability of the sample preparation and analytical procedures. These duplicate samples were generated from the crushed reject following primary crushing and were submitted to the laboratory as blind samples. The DUP samples represent an average insertion rate of approximately 0.30% (1,154 total samples).

In 2025, Aya introduced a second coarse reject duplicate sample type, recorded as DUP_REJECT. These samples comprised crushed coarse reject material randomly selected from approximately one month's sample preparation and were submitted to the laboratory as blind samples, though this has not continued into 2026. DUP_REJECT samples represent an average insertion rate of approximately 0.06% (218 samples in total). As both sample types assess the repeatability of the crushing, splitting and analytical stages of the sampling protocol, SRK has evaluated the two datasets collectively within this section.

SRK reviewed the coarse reject duplicate results for Au and Ag via scatter graphs (Figure 12‑15) and considers that, in general, the duplicate samples are in reasonable agreement (Ag R2=0.73 and 0.43 and Au R2 = 0.63 and 0.19) with their corresponding original sample grades, as demonstrated by the moderate to high statistical correlation and visual inspection of the scatter plots, other than DUP_REJECT for Au which has a R2 of 0.19. This can be explained due to the limited number of duplicate pairs, some of these elevated grades exhibit greater variability within the higher-grade mineralized intervals. In addition a small number of duplicate pairs also comprise one sample reported at, or close to, the analytical detection limit and the corresponding paired sample returning an elevated Au grade (this clearly impacts DUP_REJECT). The coarse reject duplicate graphs are also impacted by zero values in the original value, though this will not impact the R2value. These discrepancies warrant further investigation by Aya to confirm the recorded assay values and ensure that no data transcription or database errors have occurred. The coarse reject duplicate dataset also comprises fewer sample pairs than the field duplicate dataset. Consequently, individual discrepancies or outliers have a proportionally greater influence on the statistical correlation observed for the coarse reject duplicates. Overall, SRK considers the coarse reject duplicate results to demonstrate acceptable sample preparation precision for the purposes of the 2026 MRE.

Given the nature of the polymetallic sulphide mineralization, the level of variability observed is considered reasonable and does not indicate a systematic issue with the sample preparation procedures.

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Figure 12‑15    Coarse rejects (DUP and DUP_REJECT) inserted by Aya into the sample stream for Au and Ag.

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12.5.8Umpire analysis

Umpire laboratory samples have been submitted to an independent laboratory to assess the analytical accuracy and reproducibility of the primary laboratory results, and to identify any potential analytical bias between laboratories, this was undertaken in 2022, 2023 and 2025. These samples were submitted at a rate of 0.16% (626 total samples).

During the 2022 to 2024 drilling programs, umpire analyses were completed by ALS Seville, Spain. The 2025 Preliminary Economic Assessment notes that, during 2024/2025, Aya commissioned a new ALS laboratory at its Zgounder Mine, Morocco, and subsequent Boumadine umpire samples were submitted to this laboratory for analysis. It should be noted this laboratory only became operational towards the end of 2024.

SRK reviewed the available umpire laboratory results for Au and Ag via scatter plots (Figure 12‑16) and considers that, in general, the Ag (R2=0.96) umpire assays are in good agreement with the corresponding primary Afrilab analytical results, though the Au values (R2=0.57) show a poorer correlation but still within acceptable tolerances for the purposes of the 2026 MRE.

Additional variability is evident at the lower end of the Au grade distribution, where the primary Afrilab analyses commonly report values at the laboratory detection limit of 0.05 g/t Au, whereas the corresponding ALS umpire assays may report lower concentrations, possibly due to the lower reporting limits of the analytical methods employed by ALS. SRK recommends that future umpire analyses are undertaken using analytical methods with reporting limits comparable to those used by the primary laboratory wherever practicable, as differences in analytical detection limits can adversely influence the statistical correlation between the primary and umpire datasets.

In addition, the umpire duplicate dataset comprises fewer sample pairs than the field duplicate dataset and is further affected by the presence of zero values. Consequently, individual discrepancies and outliers have a proportionally greater influence on the overall statistical correlation observed for the umpire dataset. SRK recommends that Aya continues its routine umpire sampling program throughout future drilling campaigns and considers submitting representative samples from the 2024 drilling program. Aya could also consider increasing the amount of umpire samples sent to ALS.

Overall, SRK considers the umpire reject duplicate results appear to show no bias associated with the primary assay laboratory.

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Figure 12‑16    Umpire duplicate samples sent to ALS by Aya for Au and Ag

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12.6SRK Comments

SRK considers the Company's QA/QC program implemented throughout the modern exploration programs to be appropriate for the purposes of geological modelling and MRE. The routine insertion of CRMs, blank samples, field duplicates, coarse reject duplicates and umpire laboratory samples provides a comprehensive framework for monitoring analytical accuracy, precision and contamination. The CRM suite provides appropriate grade coverage for the principal analytes of economic interest (Au, Ag, Cu, Pb and Zn), and SRK's review indicates that the vast majority of CRM results fall within the expected control limits. Although a limited number of CRM failures were identified, these are primarily associated with low-grade standards containing analyte concentrations close to the analytical reporting limits and historical CRM sample ID assignment issues which were rectified by Aya.

SRK recommends that OREAS 100A, OREAS 354 and OREAS 994 are no longer included in the sample stream, as these CRMs do not have certified values for the applicable aqua regia or fire assay methods employed by Afrilab. Consequently, results for analytes without certification for the applicable analytical method are not directly comparable with the analytical methods employed by Afrilab. In some cases, certified values are available for alternative analytical methods, including four-acid digestion or classical wet chemistry, typically for CRMs containing elevated base metal concentrations. Although SRK assigned a lower level of confidence to comparisons based on these alternative certified values, no material issues were identified in the performance of these CRMs. Other CRMs certified for the applicable aqua regia digestion method provide appropriate coverage of the grade ranges encountered within the Project mineralization and, therefore, the absence of directly comparable certified values for OREAS 100A, OREAS 354 and OREAS 994 is not considered material to the outcome of the QA/QC review.

A small number of apparent sample swaps or mislabelling issues remain within the QA/QC database, resulting either from the mislabelling of original drill core samples or from samples being incorrectly assigned to other QA/QC sample types, including potentially other CRMs. This includes apparent misassignments between OREAS 234B and OREAS 236. SRK also identified two anomalous Zn CRM results reporting concentrations of 123% and 1,299%, which are considered likely to represent transcription or database errors. SRK recommends that Aya investigate these outstanding discrepancies and correct the QA/QC database where appropriate.

SRK also notes that drill core from the 2018 to 2021 drilling programs was relogged and resampled accompanied with QA/QC samples by Aya during 2023. Consequently, the QA/QC data associated with these drillholes generally reflects the 2023 resampling program, rather than the original drilling dates, and this should be considered when interpreting temporal trends within the QA/QC dataset.

The blank sample review indicates that contamination during sample preparation and laboratory analysis is generally well controlled. Field and coarse reject duplicate results demonstrate acceptable analytical precision for the style of polymetallic mineralization encountered at the Project, although a small number of database issues were identified, including zero assay values and missing Au duplicate results, which SRK recommends Aya investigate and correct prior to the next MRE update. SRK recommends that Aya incorporate routine pulp duplicate samples into future QA/QC programs to provide an additional assessment of pulverization performance and sample homogeneity.

Overall, no serious QAQC issues are evident from these results. The QAQC analyses presented are adequate and the data is of a sufficient quality to support the subsequent geological modelling, grade and tonnage estimates presented in this report, though there is opportunity for improvement.

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13.0MINERAL PROCESSING AND METALLURGICAL TESTING

13.1Introduction

Several testwork programs have been carried out at different laboratories for the Boumadine Project, with the first pilot test program carried out between 1986-1991 and the latest lab test program completed in 2024. Tested materials have included hist

orical tailings, fresh rock, and produced pyrite concentrate samples. A summary of both the historical and recent test programs is provided in Table 13-1.

Table 13-1    Summary of Testwork Programs

Testwork Program

Mineralogy

Comminution

Gravity

Flotation

Direct Cyanidation/CIL

Chlorination

Roast – CIL/Cyanidation

POX – CIL/Cyanidation

BIOX – CIL/Cyanidation

Albion – CIL/Cyanidation

MACH Reactor - Cyanidation

SODIM / BRPM SODECAT 1986-1991 (Pilot-scale)

P

URSTM 2011 (tailings sample)

P

Nicromet 2013 (tailings sample)

P

CINF 2017 (tailings sample)

P

P

P

SGS Lakefield 2018 (tailings and fresh rock samples)

P

P

P

P

P

SGS Lakefield 2019 (fresh rock sample)

P

P

P

P

P

SGS Lakefield 2022 (composited fresh rock sample)

P

P

P

P

P

P

P

P

AFRILAB 2023 Sample Campaign (fresh rock)

P

SGS Lakefield 2023-2024 (composited fresh rock and pyrite concentrate samples)

P

P

P

P

P

Mintek 2024 (pyrite concentrate sample)

P

P

BGRIMM 2025 (bulk pyrite concentrate sample)

P

P

Details of the historical testwork (1986-2017) can be found in the published Technical Report and Preliminary Economic Assessment on the Boumadine Project issued on 24 May 2019. Since the results have not been used in the development of the process design criteria, they will not be further discussed in this report. Details of the testwork done by Mintek (2024) have also not been provided since the results were not used toward the development of the process plant design and contained confidential information.

The metallurgical sampling locations for the material tested in the recent programs (2021-present) are shown by disks in the cross-sectional model renderings shown in Figure 13-1.

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Figure 13-1    Metallurgical Sampling Locations

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13.1.1SGS Lakefield Testwork Program – 2018

In 2018, SGS Lakefield conducted a metallurgical test program on fresh rock and tailings samples from the Boumadine Project. The fresh rock sample (Sample #1) was screened at 10 mesh and any oversize was crushed to pass 10 mesh. It was then homogenized and rotary split into the required test charges. A total of nine tailings samples (Sample #2, 10350, 10351, 10352, 10353, 10358, 10359, 10360 and 10361) were received, which were also homogenized (individually) and split the required test charges.

The testwork program included:

•Head analysis.

•Grindability testing (for fresh rock sample only).

•Flotation testing.

•Cyanidation testing.

•Roast-CIL testing.

•POX-CIL testing.

13.1.1.1Head Analysis

Representative aliquots of each sample were submitted for chemical analysis including whole rock analysis, ICP scanning, fire assay for Au and Ag, LECO for total sulphur and sulphides, XRF for Pb, Zn, and Cu by, and comprehensive analysis for other elements by ICP-MS and fire assay / ICP. The results for elements of interest are shown in Table 13-2.

Table 13-2    Head Analysis of Fresh Rock and Tailings Samples

Element			Units			Sample ID
						Sample #1			Sample #2			10350			10351			10352			10353			10358			10359			10360			10361
S.G.			-			4.04			3.4			2.31			2.41			2.31			3.1			-			-			-			-
Fire Assay
Au			g/t			4.29			3.20			1.79			1.34			1.58			2.16			1.75			2.10			3.26			2.52
Ag			g/t			-			-			-			-			-			-			120			165			201			136
LECO
S			%			37.0			27.8			12.4			14.9			13.0			23.3			14.4			24.1			22.8			28.7
S=			%			36.1			24.5			0.40			5.21			0.53			20.0			-			-			-			-
XRF
Pb			%			0.41			0.30			0.31			0.23			0.33			0.64			0.35			0.55			0.31			0.30
Zn			%			1.34			0.098			0.77			1.00			0.68			1.03			0.60			1.19			0.21			1.12
Cu			%			0.18			0.012			0.028			0.045			0.043			0.17			0.044			0.12			0.016			0.051

13.1.1.2Grindability Testwork

The fresh rock sample was submitted for the Bond ball mill grindability test performed at 100 mesh of grind. The Bond ball mill work index (BWi) was determined to be 10.7 kWh/t, falling in the moderately soft range of SGS’s database.

13.1.1.3Flotation Testwork

Rougher Kinetic Flotation Testing

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Open circuit batch rougher kinetic flotation tests were conducted on the Sample #1 (fresh rock) and Sample #2 (tailings).

The fresh material sample responded reasonably well to conventional lead / zinc / pyrite flotation process. Open circuit batch rougher flotation tests on the fresh rock sample resulted in 69.7% recovery of lead to a lead rougher concentrate at a grade of 7.8% Pb, and 16.7% of the gold also reported to the lead concentrate. The zinc rougher concentrate graded 36.8% Zn with 81.9% recovery at low mass pull (3.3%). The pyrite rougher circuit recovered 92% of the sulfur with high mass pull of 68.6%. The pyrite rougher concentrate sulphur grade was 50.7% and recovered 79.5% of the gold.

For the tailings sample, it was discovered that the pH of the pulped sample was very low following grinding. The quantity of lime required to raise the pH to 8.5 for flotation was very high and weight loss was observed during flotation. For the second test on Sample #2 material, the sample was washed, ground, filtered and washed again prior to flotation. The filtrate and wash solutions were analyzed and showed a high concentration of iron, indicating the presence of soluble salts which were contributing to the low pH and high lime consumption. A bulk sulphide flotation was performed on the washed tailings, producing a pyrite rougher concentrate recovering 99.6% of the sulphides and 91.3% of the gold at grades of 49.3% S and 6.1 g/t Au.

Cleaner Flotation Testing

Open circuit batch cleaner flotation tests were performed on Sample #1. The first test attempted to produce lead, zinc, and pyrite concentrates. A lead 4th cleaner concentrate grading 33.1% Pb was produced with 56.7% Pb recovery. The zinc 3rd cleaner concentrate assayed 51% Zn with 78% Zn recovery. The pyrite rougher concentrate contained 49.8% S and 3.4 g/t Au with recovery rates of 92.4% S and 73% Au. The pyrite rougher tails assayed 0.2 g/t Au. Most of the gold losses were in the lead circuit.

A second test was performed without lead flotation. A zinc concentrate was produced grading 46.4% Zn, but with only 47.6% recovery. The pyrite rougher concentrate assayed 49.8% S and 5.28 g/t Au, with 97.5% S recovery and 87.5% Au recovery. The gold loss to the zinc circuit was 11%, and the pyrite rougher tails assayed 0.22 g/t Au. 85.2% of the contained lead reported to the pyrite concentrate.

A third test was performed to produce pyrite concentrate for downstream testing (roast / CIL). A lead rougher stage was included to reduce the amount of lead in the pyrite concentrate. A zinc concentrate grading 54.2% Zn was produced with 79.5% Zn recovery. The pyrite concentrate graded 51% S and 5.18 g/t Au with sulphur recovery at 93.9% and gold recovery at 80.9%. The major gold loss was to the lead rougher concentrate at 15.5%. The pyrite concentrate was submitted for a roast / CIL test.

13.1.1.4Cyanidation Testwork

Five cyanide bottle roll tests were completed on tailings samples (Sample #2, 10350, 10351, 10352 and 10353) and two tests on pyrite concentrate samples after fine grinding to <10 microns (μm).

The tailings samples were washed before leaching. Lime and NaCN consumptions were high and further weight losses occurred during the leaching period. The normalized recoveries (based on residue and direct head grades) ranged from 25-69%.

The lime and NaCN consumptions for the pyrite concentrate samples were also high, and the overall normalized recoveries were 38-41%.

Silver recoveries were higher with the overall normalized recoveries ranging between 47-98% for the tailings samples and 59-73% for the pyrite concentrate samples.

13.1.1.5Roast-CIL Testwork

Four roasting tests were completed – two on the obtained flotation concentrate and two on the tailings composite samples. The tests were conducted as two-stage roasts, in a static muffle furnace for a total of

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four hours (two hours per stage). The sample was stirred every 15 minutes during the second stage. All sulphide sulphur was oxidized during roasting.

Each calcine was subsequently subjected to a standard cyanidation test as CIL, with 2 g/L NaCN, 10 g/L carbon for 24 hours. The RST-2 calcine (CN-9 feed) was ground in an attrition mill to <10 μm prior to cyanidation. The CIL testing results are listed in Table 13-3.

Gold extraction from the calcines ranged from 45% to 85%, whereas silver recovery ranged from 5% to 72%. The poor recovery from the fully oxidized calcine could have been due to the presence of 0.15% to 0.25% lead in the pyrite concentrates, which was known to cause problems in the roasting process.

13.1.1.6POX-CIL Testwork

Seven pressure oxidation (POX) tests were completed on the obtained flotation concentrate and tailings composite samples. The POX tests were carried out in standard Parr 2 L titanium autoclaves. The pulp density in the POX tests was dictated by the sulphide sulphur grade in the POX feed and the need to create autothermal heating conditions. The POX feed was pre-acidulated to pH 1.5 for 30 minutes by addition of sulphuric acid. After POX, the pulp was filtered, and the solids were washed. POX residues were sub-sampled for analysis and the remaining residue forwarded for CIL. POX-4 through POX-7 tests evaluated the effect of Hot Curing (HC) and Lime Boiling (LB) on subsequent lime consumption and silver recovery during CIL. The POX-CIL results are summarized in Table 13-4.

Gold recovery from the flotation concentration after POX 1 pressure oxidation (POX CIL1) was high, at 97.4%. However, lime consumption was extremely high at 321 kg/t (POX feed basis). Silver recovery was low at 49%. The results are related to the chemistry in the pressure oxidation process, with high sulphide feeds tending to produce significant amounts of basic iron sulphate and silver jarosite. The sulphate ion in basic iron sulphate does not react with limestone and can only be neutralized with lime. Therefore, copious amounts of lime were consumed during neutralization of the POX residue prior to and during cyanidation. In addition, it is well known that silver jarosite is insoluble in cyanide solution. These are well known phenomena in processing of refractory gold sulphide concentrates by pressure oxidation.

The consumption of lime by the POX 4 residue was reduced significantly to 16.8 kg/t in the POX CIL 4 test, by employing a Hot Cure stage after pressure oxidation. Silver recovery in this test, however, dropped from 49% in POX CIL 1 to just 1.2%, indicating that more silver jarosite formed during hot curing.

In the final test on the flotation concentrate, POX CIL 6, the autoclave slurry was first subjected to Hot Curing, followed by solid-liquid separation and washing of the residue. The washed solids were then subjected to a Lime Boil stage, which is designed to breakdown the silver jarosite complex. Excess lime is added to the slurry (generally 100 to 200% stoichiometric excess of the sulphate in the solids), which is then held near boiling temperature for several hours. The overall lime requirement was 60 kg/t, and recovery of both gold and silver were high after lime boiling, at approximately 98%.

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Table 13-3    Roast-CIL Results

Test ID			Feed			Temperature °C						Sulphide S %						Reagent Addition kg/t						Reagent Consumption kg/t						CIL Recovery %
						Stage 1			Stage 2			Stage 1			Stage 2			NaCN			CaO			NaCN			CaO			Au			Ag
RST-1			F7 Pyrite Con			400			700			49.6			<0.05			2.82			1.16			1.02			1.16			62.5			9.1
RST-2			240g F7 Pyrite Con +10g Pb Con			400			700									4.17			4.29			3.91			4.27			84.8			72.5
RST-3			Oxidized Tailings Comp			650			800			8.62			<0.05			2.62			1.19			1.02			1.18			68.7			7.0
RST-4			Non-Oxidized Black Tailings Comp			650			800			24.3			<0.05			3.39			1.54			1.54			1.52			44.8			4.7

Table 13-4    POX-CIL Results

Test ID			Feed			Product to CIL			Lime Boil CaO addition						Reagents (kg/t of POX feed)												CIL Recovery %
															Added						Consumed
									kg/t LB feed			kg/t POX feed			NaCN			CaO			NaCN			CaO			Au			Ag
POX CIL 1			F7 Pb Ro Con & F7 Pyrite Con			POX 1 residue			-			-			7.24			320.8			3.55			320.8			97.4			49.2
POX CIL 2			Oxidized Tailings Comp			POX 2 residue			-			-			1.96			93.2			1.00			93.2			87.7			12.1
POX CIL 3			Non-Oxidized Tailings Comp			POX 3 residue			-			-			2.21			73.7			0.61			73.6			95.4			3.5
POX CIL 4			F7 Pb Ro Conc & F7 Pyrite Conc			POX 4 HC residue			-			-			6.56			16.8			4.45			16.8			97.6			1.2
POX CIL 5			Non-Oxidized Tailings Comp			POX 5 HC residue			-			-			1.49			2.4			0.37			2.2			43.7			65.7
POX CIL 6			F7 Pb Ro Conc & F7 Pyrite Conc			POX 6 HC-LB residue			263			59.2			2.11			0.6			0.90			0.6			98.2			97.6
POX CIL 7			Non-Oxidized Tailings Comp			POX 7 HC-LB residue			40.1			23.8			1.39			0.4			0.25			0.3			71.6			81.4

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13.1.2SGS Lakefield Testwork Program - 2019

In 2019, further metallurgical testwork was carried out on the fresh rock sample that was used in previous 2018 program. The sample was stored in the freezer and therefore was considered with the same characteristics to samples tested in the previous program.

Sequential flotation was carried out on the ground sample to produce a pyrite concentrate for gold recovery testwork, including baseline cyanidation, POX, and Albion Process™. The Albion Process™, developed by Xstrata Pty Ltd. (Glencore Technology Pty Limited) is a combination of fine grinding and exothermic oxidative leaching without the need for pressure treatment or external heating.

Open circuit batch rougher flotation tests resulted in a lead rougher concentrate with a grade of 10.4% Pb at a recovery of 67.0%, along with 12.9% of the gold and 43.7% of the silver, which also reported to the lead concentrate of grades of 19.8 g/t and 6,906 g/t, respectively. The zinc rougher concentrate graded 51.1% Zn at 68.8% recovery with a low mass pull (1.9%). The pyrite rougher circuit recovered 93.9% of the sulphur to the concentrate, with a high mass pull of 70%.

A full chemical head analysis and a semi-quantitative XRD analysis were conducted on the pyrite concentrate. The concentrate had a sulphur grade of 51.0% with 4.96 g/t Au and 310 g/t Ag, and the concentrate sample was found to be composed of 94% pyrite with just over 5% quartz and less than 1% each of arsenopyrite and chalcopyrite.

A bottle roll cyanidation test was conducted on the pyrite concentrate to serve as a baseline for comparing the gold recoveries after processing the pyrite concentrate by either the Albion or POX processes. The concentrate was first attrition-ground (250 g at 50% solids for 30 minutes), resulting in a product D80 of 4.47 µm (determined via Malvern analysis). The CIL conditions were 5.0 g/L NaCN with oxygen sparging at 40% solids with 15 g/L carbon for 24 hours.

The Albion Process™ testing at SGS Lakefield is accredited by Glencore Technology. The attrition-ground solids were pulped to 10%, heated to 95 °C and the slurry was first acidified using sulphuric acid to a pH of 4.5 until the pH started to fall naturally. Oxygen was sparged into the pulp at ~1 L/min and alkali addition in the form of 20% solids limestone slurry (CaCO3) was added via a pH controller to control the pH at 5.5. The test ran for a period of ~52 hours with constant monitoring of pH, ORP and limestone additions were recorded. Hot water was added to the pulp throughout the leach to make up for any evaporative losses. The residue from Albion leaching was then subjected to CIL, under the same conditions as for the baseline test.

With the relatively low level of sulphide oxidation obtained in the first Albion leach test (Neutral Albion-1R), a repeat test was carried out. In the repeat Albion test (Neutral Albion-1R), the pyrite concentrates leach residence time was increased from 52 to 72 hours and the oxygen sparging rate was increased from 1 L/min to 1.7 L/min. Initial acid addition to kick-start oxidation was 2 kg/t lower than in the first test at 21 kg/t H2SO4. Alkali addition of 1,209 kg/t CaCO3 was almost three times the amount used in the first test (440 kg/t).

Three identical POX tests were conducted, and the products were combined for downstream testing (hot cure, lime boil and CIL). The pulp was pre-acidulated for 30 minutes at pH 1.5 with sulphuric acid in 2 L titanium autoclave vessels. After 30 minutes, the autoclave was sealed, pressure tested and heated to target temperature. The tests were run at 225 °C for 60 minutes, with 690 kPa (~100 psi) of oxygen over pressure applied. The three POX pulps were combined and heated to 95°C for four hours in a 4 L Pyrex vessel with a lid and overhead mechanical agitation (Hot Curing process, to re-dissolve basic iron sulphate to ferric sulphate in solution). The washed hot cure residue was pulped to 10% solids and boiled at 95°C for six hours at a reagent addition of 35.8 kg/t POX feed (Lime Boiling process, to break down the silver jarosite complex). After six hours, the pulp was forwarded for CIL testing.

The CIL testing results of the three gold recovery processes are summarized in Table 13-5. The baseline test had high addition and consumption rates for NaCN and demonstrated good silver recovery at 84%, but poor recovery of 17%. CIL on the Albion leached product (NA-1, obtained from 52 hours Albion leaching) resulted in

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slightly higher gold recovery to 45% with similar silver extraction (85%) as the baseline. The repeated Albion test (NA-2, 72 hours) resulted in CIL leaching recovery of 81% for gold and 92.3% for silver. It was anticipated that with further optimization of the Albion conditions, further sulphide oxidation can likely be achieved and gold extraction from the residue increased even further. Pressure Oxidation of the pyrite concentrate, followed by hot curing and lime boiling, provided the best extraction of the three processes – with extractions of 98% and 96%, respectively.

Table 13-5    CIL Results Summary

Test			Reagents (kg/t of CIL feed)												Recovery %						CIL Residue g/t
			Added						Consumed
			NaCN			CaO			NaCN			CaO			Au			Ag			Au			Ag
CIL-1 Baseline			24.9			1.83			22.3			1.60			17.3			83.9			4.07			48
CIL-2 Neutral Albion-1			3.85			11.7			3.52			11.6			45.4			84.8			1.65			29.4
CIL-4 Neutral Albion-1R			5.46			6.79			5.11			6.79			81.0			92.3			0.37			8.6
CIL-3 POX-HC-LB*			2.23			3.98			1.52			3.98			98.2			96.6			0.41			47.6

*35.8 kg/t CaO added during LB prior to CIL-3.

13.1.3SGS Lakefield Testwork Program - 2022

The 2022 test program by SGS Lakefield was carried out on a composite of 75 samples received from the Boumadine deposit (called the Main Composite). The samples were crushed, blended and split to portions for chemical and mineralogy analysis, comminution testing (BWi), and metallurgical testing including:

•Gravity separation.

•Flotation (open circuit and locked cycle).

•Direct CIL cyanidation of pyrite flotation concentrate.

•BIOX-CIL.

•Albion Leach™-CIL.

•Roast-CIL.

13.1.3.1Head Analysis

The chemical analysis included gold and silver by fire assay, total organic carbon, total carbon, sulphide sulphur, mercury, lead, zinc, and a multi-element ICP scan. The results for the elements of significant interest are shown in Table 13-6.

Table 13-6    Composite Sample Head Analysis

Element			Au g/t			Ag g/t			As g/t			Cu %			S= %			Pb %			Zn %
Assay			3.33			113			834			1.94			26.5			1.00			2.25

13.1.3.2Mineralogy

X-Ray Diffraction

The XRD results provided the overall mineral composition of the composite sample as summarized in Table 13-7. Slightly more than 40% of the sample weight were sulphides with the most common being pyrite.

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Table 13-7    Composite Sample Mineralogy

Mineral			% Weight
Quartz			34.6
Pyrite			32.6
Muscovite			17.1
Chlorite			3.3
Sphalerite			3.2
Gypsum			2.9
Arsenopyrite			2.8
Biotite			1.7
Galena			0.9
Chalcopyrite			0.9
Total			100

TIMA-X

A portion of the sample was stage-crushed, targeting a P80 of 75 µm but obtaining a true P80 of 220 µm. The sample was screened into +53 µm and -53 µm fractions. The modal mineralogy, liberation and association characteristics of sulphides, and gold and silver deportment of the samples were characterized by Tescan Integrated Mineral Analyzer (TIMA-X) technology using the High-Resolution Mapping method. The results for both size fractions were combined to represent the composite sample.

Modal Mineralogy

The sulphides in the composite sample were found to consist of mainly of pyrite (55.5%), sphalerite (4.1%), arsenopyrite (3.2%), galena (1.4%), and trace amounts of chalcopyrite (0.4%), pyrrhotite (0.1%), and tetrahedrite (0.06%). The remainder of the sample was found to be composed of quartz (21.7%), micas/chlorite/clays (10.6%), and trace amounts (<1% each) of other minerals.

A discrepancy was observed between the quartz and pyrite grades obtained by the XRD and TIMA analyses, which was attributed to potential gravity settling during sample preparation.

Sulphide Liberation and Association

At the grind size P80 of 220 µm, the liberation and mineral association of the key minerals of interest were as follows:

•A high amount of the contained pyrite was found to be free and liberated (approximately 94%). The increase of pyrite liberation from the +53 µm to -53 µm size fraction was minimal at 2%.

•Approximately 77% of the contained arsenopyrite was free and liberated. The increase of arsenopyrite liberation between the +53 μm to -53 μm size fractions was significant at 24%.

•Approximately 78% of the contained sphalerite was free and liberated. The increase of sphalerite liberation between the +53 μm to -53 μm size fractions was significant at 47%.

•Approximately 65% of the contained galena was free and liberated. The increase of galena liberation between the +53 μm to -53 μm size fractions was significant at 60%.

•The remainder of the arsenopyrite, sphalerite and galena occurred as complex particles and middlings with pyrite and quartz / feldspars.

Based on individual particle liberation and grade, mineralogically limited grade-recovery analyses were carried out to indicate theoretical maximum achievable grade by recovery of the sulphide minerals. The analysis projected:

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•Galena grades between 100% and 58% for recoveries of between 54% and 87%.

•Sphalerite grades between >99% and 63% for recoveries of between 69% and 96%.

•Arsenopyrite grades between >99% and 72% for recoveries of between 56% and 96%.

Gold and Silver Mineralogy

69% of the contained gold in the composite sample was determined to be native gold and 30% was electrum. The gold grains were generally very fine (<15 µm) with 78% of the gold grains below <3 µm, accounting for 23% of the gold by mass. The gold minerals were poorly liberated, with pure, free and liberated grains accounting for only 12.4% of the total. Non-liberated grains occurred as middlings with pyrite (45.8%), and sphalerite (1.8%), and complex particles (40.0%). These phenomena suggested that achieving high gold concentrations by physical means would be unrealistic, and that leaching extraction would need to follow a high degree of pyrite oxidation.

Silver was found to be widely distributed as native silver (20%) with the remainder occurring as acanthine, electrum, tetrahedrite, polybasite, freibergite, jalpaite, native gold, and others. Only 1.6% of the silver minerals were free, pure and liberated, with the remaining non-liberated particles occurring as middlings with pyrite (25.8%) and quartz / feldspars (4.1%), and predominantly as complex particles (67.4%).

13.1.3.3Comminution Testing

The Main Composite sample was submitted for the Bond ball mill grindability test performed at a 150 mesh of grind. The measured Bond ball mill work index (BWi) was 15.4 kWh/t, falling in the moderately hard range when compared to SGS’s database.

13.1.3.4Gravity Separation Testing

1.8 kg of the Main Composite sample was used for the test feed. The feed was ground in a laboratory rod mill to a P80 of 78 μm. The mill discharge was passed through a Knelson MD-3 Concentrator, collecting a concentrate and producing tailings. The Knelson concentrate was further upgraded on a Mozley mineral separator. The final concentrate represented 0.1% of the feed mass and graded 312 g/t Au and 747 g/t Ag. The combined Knelson and Mozley tailings assayed 1.55 g/t Au and 117 g/t Ag. The recovery to the final Mozley concentrate was 19% Au and 0.7% Ag. The results indicated that gravity separation would add little value to the flowsheet.

13.1.3.5Flotation Testwork

Open circuit rougher and cleaner tests were completed on the Main Composite sample, attempting to produce a saleable lead / gold / silver concentrate and a pyrite concentrate with the maximum amount of gold and silver, along with a zinc concentrate that would be floated prior to pyrite flotation. Locked cycle testing was also completed to determine the effect of recirculating middling products, to test the robustness of the conditions and to produce a pyrite concentrate for downstream testing.

Rougher Flotation Tests

The baseline lead-zinc-pyrite kinetic test was completed at a primary grind P80 of 58 μm. The reagents and grind size were varied in the subsequent tests. These test conditions produced a lead-gold-silver concentrate grading 9.39% Pb, 3.47% Zn, 1.60 % As, 30.8% S, 12.4 g/t Au, and 678 g/t Ag in the baseline test. The concentrate recovered 87% of lead, 32.8% of gold, and 53.9% of Ag. However, 14.8% of the Zn and 13% of the As also reported to the lead rougher concentrate.

Further tests were conducted with the objective of increasing the grade of the lead-gold-silver concentrate and a pyrite concentrate containing the remainder of the gold and silver. The tests investigated different reagent schemes and grind sizes. The main objective of these tests was to reduce the amount of zinc reporting to the

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lead concentrate. Various zinc / sulphide depressants, in addition to ZnSO4 and NaCN, were tested, such as sodium sulphide, sodium sulphite, and sodium metabisulphite. None of these alternatives significantly reduced the level of zinc in the rougher concentrate. The zinc recovered to the lead concentrates ranged from14% to 25%. The combined lead + pyrite concentrates all achieved recoveries of greater than 90% gold and 85-87% silver.

Two flotation tests were performed at finer grind sizes of P80 of 37 μm. The finer grind did not appear to have any effect on overall grades or recoveries, although reagent additions also differed from the coarser ground tests.

Bulk Sulphide Flotation Test

A single bulk sulphide flotation test was conducted a neutral pH, producing a concentrate with 1.53% Pb, 3.52% Zn, 1.79% As, 44.1% S, 5.16 g/t Au and 185 g/t Ag. The recovery of all elements was very high (>96%), but the concentrate mass pull was also very high at 58.7%.

Cleaner Flotation Tests

Four open circuit batch cleaner flotation tests were performed on the Main Composite sample, trialling various circuit arrangements, grind sizes and flotation reagents conditions. The results showed that adding sodium cyanide to the lead flotation circuit helped to depress sphalerite and pyrite flotation, decreasing the amount of zinc in the lead concentrate by 50%. Implementing a zinc regrind helped improve the zinc concentrate grade slightly but the recovery increased more significantly from 59.7% to 74.8%. Implementing a finer lead grind produced a higher-grade concentrate with lower zinc content but reduced overall recovery. These trials were able to achieve sufficient lead recovery and zinc grade but had high grades of gold and silver in the lead concentrate, lead grades below the saleable target, and low zinc recovery. The results suggested that to achieve concentrate grades of 50% or more for both lead and zinc, low recoveries would be expected.

Locked Cycle Flotation

A locked cycle test was completed with the goals of assessing overall metallurgical performance with the incorporation of internal recycles and producing concentrate for downstream testing. The test was completed over six cycles with 2 kg of feed per cycle. Products were recirculated in a typical counter-current manner. The lead and zinc first cleaner tailings were advanced to the next circuit roughing stage. The tested flowsheet, including the recirculation of middling streams, is shown in Figure 13-2.

Figure 13-2    Locked Cycle Test Flowsheet

The conditions for the locked cycle test were based on the results from the batch tests, with minor adjustments in the reagent dosages. The primary grind P80 target was 58 μm, lead regrind D80 was 18 μm, and zinc regrind D80 was 17 μm. The test conditions are summarized in Table 13-8.

The test generally exhibited satisfactory stability except for zinc, where the stability was lower than the other elements, suggesting that steady state may not have been achieved in the zinc circuit. The results of the locked cycle tests based on the out-products from the final 3 cycles are presented in Table 13-9.

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Table 13-8    Locked Cycle Test Conditions

Pb Circuit			Reagents Dosed (g/t)															Time (min)
			Lime			Zn(SO4).7H2O			NaCN			Aero3418A			MIBC			Grind			Cond.
Grind			-			300			100			5			-			70
Pb Rougher 1			-			-			-			-			10						1
Pb Rougher 2			-			15			5			5			5						1
Pb Rougher 3			-			15			5			5			5						1
Pb Regrind			-			30			10			-			-			15
Pb 1st Cleaner			-			-			-			5			5						1
Pb 1st Cl Scav			-			-			-			2.5			5						1
Pb 2nd Cleaner			-			-			-			-			5						1
Pb 3rd Cleaner			-			-			-			-			5						1
Total			0			360			120			22.5			40						7
Zn Circuit (Pb Ro + 1st Cl Sc Tails)			Reagents Dosed (g/t)												Time (min)
			Lime			Cu(SO4).5H2O (10% w/w)			SIPX (1% w/w)			MIBC			Grind			Cond.			Froth
Zn Conditioner 1			1,400			-			-			-						3
Zn Conditioner 2			-			225			-			-						3
			265			-			-			-						2
Zn Rougher 1			-			-			10			10						1			1
Zn Rougher 2			165			-			10			7.5						1			2
Zn Regrind			250			-			5			-			5
Zn 1st Cleaner			60			-			-			7.5						1			2
Zn 1st Cl Scav			-			-			5			7.5						1			1
Zn 2nd Cleaner			50			-			-			7.5						1			1.5
Zn 3rd Cleaner			140			-			2.5			5						1			1.5
Total			930			225			32.5			45						11			9
Sulphide Circuit (Zn Ro + Zn 1st Cl Tails)			Reagents Dosed (g/t)									Time (min)						pH			Eh
			PAX			MIBC			H2SO4			Cond.			Froth
Sulphide Conditioner			-			-			1,000			1						8.0			100
Sulphide Rougher 1			100			10			-			1			2			8.2			75
Sulphide Rougher 2			100			7.5			-			1			4			8.4			50
Sulphide Rougher 3			100			10			-			1			8			8.5			50

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Table 13-9    Locked Cycle Test Results Summary

Product			Weight %			Assays																					% Distribution
						Cu %			Pb %			Zn %			As %			S %			Au g/t			Ag g/t			Cu			Pb			Zn			As			S			Au			Ag
Pb Conc			3.0			4.81			26.7			4.35			1.67			35.0			26.6			1,923			75.3			84.5			7.1			4.7			4.1			26.5			53.1
Zn Conc			2.3			0.37			0.48			58.1			0.27			32.9			1.22			166			4.3			1.2			72.0			0.6			2.9			0.9			3.5
Pyrite Conc			54.9			0.061			0.18			0.65			1.69			42.4			3.80			78.5			17.4			10.6			19.5			86.8			89.9			69			39.6
Pyrite Tails			39.9			0.015			0.089			0.063			0.21			2.02			0.27			10.5			3.0			3.7			1.4			7.9			3.1			3.6			3.8
Head (calc)			100			0.19			0.95			1.83			1.07			25.9			3.02			109			100			100			100			100			100			100			100
Head (direct)			-			0.19			1.00			2.25			0.083			-			3.33			113			-			-			-			-			-			-			-

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The following observations were made from the locked cycle test:

The final lead concentrate had a good lead recovery of 84.5%, but low grade of 26.7% Pb. The main diluents were copper and zinc.

•The final zinc concentrate grade was good at 58.1% Zn in 2.3% of the total mass, but the recovery was modest at 72%.

•The pyrite concentrate recovered 69% of the gold, 39.6% of the silver, 89.9% of the S, and 86.8% of the arsenic, in a high mass pull of 54.9%.

•Gold and silver were mainly recovered to the lead concentrate, with 26.5% of the gold and 53.1% of the silver reporting to the lead concentrate.

13.1.3.6Pyrite Concentrate Testing

The pyrite concentrates from cycles the six were combined and blended, then split into small charges for direct CIL, Albion, BIOX, and roasting testwork. The P80 size of the concentrate composite was determined to be 42 µm, and the composition of elements of interest is provided in Table 13-10.

Table 13-10    Pyrite Concentrate Analysis

			Au g/t			Ag g/t			Pb %			Zn %			Fe %			As %			S %			S= %
Assay			3.83			76.2			0.27			0.45			35.8			1.72			42.8			41.4

Direct CIL Testing

Two cyanidation tests were completed on the pyrite concentrate at two grind sizes: one at the “as received” grind size of P80 42 μm and the other reground to a D80 of 6.3 μm. The other test conditions were held constant across both trials. The results of the tests are summarized in Table 13-11.

Table 13-11    Direct CIL Extraction Results

Test No.			P80 µm			Reagent Added kg/t CN Feed						Reagent Cons. kg/t of CN Feed						Au Extraction %			Ag Extraction %
						NaCN			CaO			NaCN			CaO
CIL 1			42			20.3			4.25			4.39			3.91			9.5			36.4
CIL 2			6.3			28.5			15.4			14.5			15.3			22.4			80.2

Gold extraction was improved with fine grinding but was still very low for both trials. Better extractions were seen for silver, reaching 80% for the finely ground sample. The reagent consumptions were high and increased significantly with fine grinding.

Albion Testing

Glencore recommended Albion leach parameters that were in line with their standard neutral leach protocol. The Albion oxidation test conditions and results are summarized in Table 13-12.

Table 13-12    Pyrite Concentrate Albion Testing Results

Test No.			Retention Time h			P80 µm			Limestone Addition kg/t			Sulphide Oxidation %
NAL-1			96			6.3			1,647			96.3
NAL-2			78			10			1,261			74.7

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Sulphide oxidation was high (96.3%) for the test at the finer grind size P80 of 6.3 μm and longer retention time but dropped significantly to 74.7% when the grind was coarsened to 10 μm and the retention time decreased to 78 hours.

The Albion residues were washed before they were subjected to cyanide leaching testing. The CIL test results are provided in Table 13-13.

Table 13-13    CIL Leaching Results for Albion Residues

Sample			Reagent Added kg/t CN Feed						Reagent Cons kg/t of CN Feed						Au Extraction %			Ag Extraction %
			NaCN			CaO			NaCN			CaO
NAL-1 Residue			7.9			2.5			3.2			2.4			94.4			94.8
NAL-2 Residue			8.2			3.3			2.3			3.3			84.9			91.3

Gold and silver extractions were greatly improved after the Albion pre-treatment. The test at 6.3 μm grind size achieved gold and silver extractions of 94.4% and 94.8%, respectively. The gold extraction decreased significantly with the coarser grind and shorter retention time. Reagent consumptions for both trials were reduced in comparison with the direct CIL tests.

Bacterial Oxidation (BIOX) Testing

The inoculum build up was performed in preparation of the bioleach tests. For each of the tests, the pyrite concentrate sample was slurried with the bacteria, with conditions maintained for optimum bacterial activity (heated to 40°C in an acidic solution, targeting ~10.5% pulp density). After 18, 26, or 35 days, the pulps were filtered, and the filtrate was collected and measured. A filtrate sample was submitted for analysis and the rest was retained. The filter cake was repulped with water, agitated and re-filtered. The filter cake was then washed three times, targeting one displacement volume of water each time a representative subsample of the washed filter cake was submitted for chemical analysis. The remainder of the filter cake was forwarded to cyanidation. The bio-oxidation liquor would have to be neutralized prior to disposal, but no neutralization testwork was conducted at this stage. The results of bio-oxidation are provided in Table 13-14.

Table 13-14    Bio-Oxidation Testing Results

Test No.			Days			Sulphide Oxidation %
BAT-2			18			60.6
BAT-3			26			88.0
BAT-1			35			92.8

Sulphide oxidation improved with increased retention time, ranging from 60.6% for the 18-day test to 92.8% for the 35-day test. It should be noted that a 35-day residence time in a batch test translates to an ~5-day retention in a continuous process.

The results of CIL on the bio-oxidation residue are provided in Table 13-15.

Table 13-15    Bio-Oxidation Residue CIL Results

Sample			Test No.			Reagent Added kg/t CN Feed						Reagent Cons. kg/t of CN Feed)						Au Extraction %			Ag Extraction %
						NaCN			CaO			NaCN			CaO
BAT-2 Residue			CIL 4			17.7			9.0			7.9			8.8			75.2			75.5
BAT-3 Residue			CIL6			15.3			11.4			8.8			11.2			82.2			40.1
BAT-1 Residue			CIL 7			23.6			16.2			23.4			16.2			85.1			37.0

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The CIL gold extractions improved from 75 to 85% with the longer oxidation period (i.e. higher sulphide oxidation), however silver extractions decreased drastically. This result is most likely due to jarosite formation and may be overcome with the inclusion of a lime boil stage.

Reagent consumption in the CIL tests was very high. High cyanide consumption in commercial bacterial leaching plants is generally associated with the formation of elemental sulphur as an intermediate sulphide oxidation product in the BIOX process, and the reaction between cyanide and sulphur to produce thiocyanate ions in solution.

Roast-CIL Testing

Five roasting tests were completed on the pyrite concentrate. The tests were conducted in a static muffle furnace under various roasting conditions (changing the number of stages, temperature and/or aeration). The samples were rabbled every 15 minutes. The results are summarized in Table 13-16.

Table 13-16    Roast Testing Results

Test			Roast Conditions			Sulphide Oxidation %
RST-1			1st stage 400°C and 2nd stage 700°C			99.9
RST-2			1st stage 400°C and 2nd stage 600°C			97.6
RST-3			1st stage 400°C, 2nd stage 600°C with 1L/min oxygen, 3rd stage 600°C with 2L/min oxygen			99.8
RST-6			1-stage 500°C			89.3
RST-7			1-stage 600°C			95.1

The results showed that oxidation improved with more than one roasting stage with higher temperatures. While oxidation was high with aeration, only one trial was conducted so it is difficult to attribute the results to aeration versus the longer retention time and temperature.

The roasted product from each trial was then split in half, with one half left as-is, and the other half reground in a lab attrition mill to a D80 of <10 μm. The calcines were then subjected to standard CIL cyanidation tests. Results of the CIL tests are provided in Table 13-17.

Table 13-17    Roast Residue CIL Results

Sample			P80 µm			Reagent Added kg/t CN Feed						Reagent Cons. kg/t of CN Feed						Au Extraction %			Ag Extraction %
						NaCN			CaO			NaCN			CaO
RST 1 Calcine			46			21.9			12.0			6.5			12.0			61.1			19.9
			7.4			24.0			6.9			4.8			6.6			71.1			44.9
RST 2 Calcine			41.5			24.1			-			13.3			-			55.6			65.9
			6.7			31.4			12.2			12.7			12.2			65.9			60.3
RST 3 Calcine			44.3			27.3			8.6			10.3			8.6			59.3			32.5
			7.5			27.3			5.9			5.1			5.5			66.7			51.6
RST 6 Calcine			51.3			29.1			34.7			13.6			34.7			52.5			51.8
			8.1			37.2			14.6			22.5			14.6			63.9			61.8
RST 7 Calcine			36.5			30.5			13.5			13.9			13.5			56.4			29.6
			5.9			26.8			9.3			11.8			9.1			66.7			50.4

Gold extraction from the calcines was poor, ranging from 55.6% to 71.1%, whereas silver extraction ranged from 19.9% to 60.3%. SGS stated that they were uncertain why gold and silver extractions were so poor, even after near complete oxidation of the sulphides. The formation of an impervious molten phase during the roasting process was suggested as a possibility.

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13.1.4SGS Testwork Program – 2023-2024

In 2024, more testing was conducted by SGS on the Main Composite and the pyrite concentrate composite produced during lock-cycle testing obtained from the previous 2022 program. The main objective of the 2024 program was to investigate methods of separating the arsenopyrite from the pyrite in the sulphide concentrate prior to oxidative treatment. The idea was that improving the gold grade and Au:S ratio would enhance the economics of sulphide oxidation by separating the higher gold grade / lower sulphide grade arsenopyrite from the pyrite in the pyrite concentrate. By treating the two streams separately, favorable conditions for each could be discerned to produce a more optimized and economic process. Several methods were tested, including gravity separation and Jameson Cell flotation.

After none of the tested methods proved suitable, the 2024 program was then re-focused on the pre-oxidation of the concentrate prior to cyanide leaching. A locked cycle test was completed on the Main Composite to produce a lead concentrate for mineralogy and a pyrite concentrate for pressure oxidation and roasting. The oxidative tests were performed to verify the recoveries achieved in previous programs between 2018-2022.

13.1.4.1Mineralogy Testing

The pyrite concentrate from the 2022 locked cycle tests had been submitted for chemical analysis at that time but was submitted for bulk mineralogy analysis as part of the 2024 testwork. The main sulphide in the sample was pyrite with trace pyrrhotite, followed by minor arsenopyrite, and sphalerite. Quartz and muscovite were the main silicates. The liberation of Fe-sulphides (pyrite and pyrrhotite) was very good at 98%, and that of arsenopyrite was 87%. Approximately 7% of the arsenopyrite was associated with Fe-sulphides. The P80 of Fe-sulphides was 41 μm; that of arsenopyrite was 38 μm.

Eighty-three gold grains were identified, ranging in size from <3 to10 μm. The gold minerals were associated mainly with pyrite and arsenopyrite and they occurred as poorly exposed attachments (47%) and locked inclusions (53%) in the sulphides. This indicated that very fine regrinding would be required to liberate or expose the gold minerals associated with the sulphides. The submicroscopic gold would not be amenable to cyanidation, while the poorly exposed gold attachments would be theoretically amenable to cyanidation.

D-SIMS analysis of pyrite yielded an average gold concentration of 1.62 ppm, and of arsenopyrite 13.5 ppm. The total amount of gold (both microscopic and submicroscopic) associated with Fe-sulphides accounted for 63%, arsenopyrite 14%, and both Fe-sulphides and arsenopyrite 18%. More than 95% of the total microscopic and submicroscopic gold was associated with Fe-sulphides and arsenopyrite. Therefore, recovery of the sulphides will also yield the maximum recovery of gold.

A total of 474 silver grains were identified, including acanthite (argentite), native silver, Ag-Cu-Sb sulphides / sulfosalts, and hessite. The silver was not liberated and mainly associated with iron sulphides (83.6%), arsenopyrite (3.5%), complex particles (10.9%), and with iron sulphides and arsenopyrite (1.8%). Approximately 17% of the silver minerals occurred as inclusions while the remainder (83%) were variably exposed. Silver minerals were fine and ranged in size from <3 to 15 µm.

13.1.4.2Gravity Separation Testing

Two 91 g aliquots of the pyrite concentrate were used to determine the potential for upgrading the gold and reducing the arsenic grade using a Mozley Mineral Separator table. One of the aliquots was ground in a lab attrition mill to a D80 grind size of 7 µm while the other was tested without a regrind. A Mozley concentrate and three Mozley middling samples, along with the Mozley tailings, were collected, filtered, dried, and assayed for Au, Ag, and As. The gold and arsenic recovery were calculated, and their relationship was plotted. It was observed that gold followed the arsenopyrite, indicating no separation of arsenic was achieved.

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13.1.4.3Flotation Testing

Open Circuit Batch Cleaner Flotation

A 2 kg test charge of the Main Composite was ground in a laboratory rod mill at 60% solids for 70 minutes obtaining a P80 particle size of 58 µm. The mill discharge was subjected to lead rougher flotation followed by zinc rougher flotation before producing a pyrite rougher concentrate from the zinc rougher tailings, using conditions established in the 2022 testwork campaign. The pyrite rougher concentrate was reground to 37 µm in a laboratory rod mill with ceramic media. The reground concentrate was cleaned in three stages and with a first cleaner scavenger. The flowsheet of the open circuit batch cleaner flotation test is presented in Figure 13-3.

Figure 13-3    Cleaner Flotation Flowsheet for Separating Arsenopyrite / Pyrite

Similar to the gravity separation tests, the Au:S ratio was not improved. While the pyrite 3rd cleaner concentrate gold grade increased from 1.41 g/t Au to 4.34 g/t Au, the sulphur grade increased even more, from 9.8% to 52.9%. Significant gold losses were observed in the cleaning stages.

Jameson Cell Dilution Test

A bench flotation test was performed to evaluate the potential of a Jameson flotation cell to upgrade gold in a cleaner application. This test followed the bench dilution test procedures supplied by Glencore Technologies in which dilute flotation conditions are used to simulate the effects of wash water on the system. A sequential test was performed to remove the lead and zinc first, then the tailings were used to evaluate the pyrite cleaner circuit.

The results indicated no benefit to dilution cleaning of the pyrite concentrate, but there was a slight improvement comparing to cleaner flotation.

Arsenopyrite-Pyrite Separation Testing

A 4 kg test charge of the Main Composite was ground to a P80 of 58 µm in a laboratory rod mill. The mill discharge was subjected to sequential lead, zinc and pyrite rougher flotation to produce enough pyrite concentrate for several tests that would investigate separating arsenopyrite from the pyrite. The obtained pyrite concentrate was split into four equal aliquots for the testwork.

Various reagent combinations were examined during the tests. The most promising results were observed using a simple combination of lime, copper sulphate and xanthate at a high pH, with 80% of the arsenic reporting to the concentrate. However, there was no significant improvement in the gold grades in the concentrate in these tests, and there was a direct correlation between gold and sulphur recoveries with R2 values very close to 1.

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Locked Cycle Test

A single 6-cycle locked cycle test was completed to produce pyrite concentrate for POX and roasting tests, using the same flowsheet as the 2022 locked cycle test. The same test conditions from the 2022 locked-cycle test were used, including residence times and reagent additions (all conditions are shown in Table 13-7). The only variance was lime addition, which was added to the zinc circuit as required to maintain the target pH. The results are summarized in Table 13-18.

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Table 13-18    Locked Cycle Test Results Summary

Product			Weight %			Assays																		% Distribution
						Cu %			Pb %			Zn %			S %			Au g/t			Ag g/t			Cu			Pb			Zn			S			Au			Ag
Pb Conc			2.1			6.02			32.5			4.36			30.3			32.0			2,244			71.4			79.4			4.9			2.4			21.0			47.7
Zn Conc			2.6			0.41			0.80			56.6			32.6			1.35			220			6.0			2.4			77.4			3.2			1.1			5.7
Pyrite Conc			52.1			0.07			0.22			0.59			46.2			4.61			83.5			20.2			13.3			16.2			90.0			73.7			43.2
Pyrite Tails			43.2			0.01			0.1			0.07			2.73			0.31			7.78			2.4			4.9			1.5			4.4			4.1			3.3
Head (calc)			100.0			0.18			0.88			1.91			26.8			3.25			101			100.0			100.0			100.0			100.0			100.0			100.0
Head (direct)			-			0.19			1.00			2.25			26.5*			3.33			113			-			-			-			-			-			-

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The results were similar to results from the previous 2022 program by SGS (shown in Table 13-8), further validating the test conditions. The gold, silver and sulphur grades and recoveries were slightly higher, and the gold and sulphur ratios in the pyrite concentrates were essentially the same (0.09 and 0.1, respectively).

With a significant silver grade, the lead concentrate from the third cleaner was submitted for silver mineralogy. The analysis found that the silver was fine-grained with most particles in tetrahedrite form, ranging widely from 0.6 to 31% w/w Ag.

13.1.4.4Pyrite Oxidative Treatment Testing

To verify the previous test results, one roast test and one POX test were conducted on the locked-cycle pyrite concentrate samples.

Roast-CIL Testing

A total of 500 g of concentrate was roasted in two 250 g batches as a two-stage roasting process, in a static muffle furnace. Each stage was two hours, and the sample was rabbled every 15 minutes. The first stage was conducted at 400°C and the second was at 700°C. Neither stage was aerated. Both batches were able to achieve sulphide oxidation of 99.9%.

After the roast, the two calcines were blended and split in half. One half was left as-is and the other half was reground in a lab attrition mill to a D80 particle size of 4 µm. The calcines were then subjected to a standard cyanidation / CIL test. The results are summarized in Table 13-19.

Table 13-19    Roast Residue CIL Results

Sample			P80 µ			Test No.			Reagent Added kg/t CN Feed						Reagent Cons. kg/t of CN Feed						Au Extraction %			Ag Extraction %
									NaCN			CaO			NaCN			CaO
Roast Calcine			49			CIL 2			5.84			5.97			0.62			5.97			73.0			11.3
			4			CIL 3			7.6			5.39			2.81			5.39			80.4			30.2

Gold recovery was 80% from the finely ground calcine by CIL and 73% from the as-received calcine. Silver recoveries were lower at 11.3% and 30.2% for the coarse and fine sizes, respectively. The comparative test from the previous program achieved a gold recovery of 71.1% and silver recovery of 44.9% after grinding to 7.4 µm (shown in Table 13-16).

It was also observed that the addition and consumption of NaCN and CaO for the program were significantly lower than the previous program (results shown in Table 13.2.16). There was no explanation on this discrepancy in the SGS report and requires further investigation. The lesser quantity of reagent added and consumed could have led to the significant drop of silver recoveries observed in the 2024 test program.

POX-HC-LB-CIL Testing

Due to the high sulphide content in the feed, the POX test was conducted at 8% w/w pulp density in DI water to simulate autothermal heating conditions in the autoclave. The pulp was pre-acidulated for 30 minutes at a pH of 1.5, and then it was sealed and heated to the target temperature of 225°C. The test was run for 690 kPa of oxygen over-pressure applied.

After the pressure oxidation test, the pulp was subjected to hot curing to re-dissolve the basic iron sulphate to reduce the lime consumption in the subsequent CIL step. The hot curing process involved heating the pulp to 95°C for four hours in the autoclave. An 80% weight loss was observed when the iron sulphates were re-dissolved, upgrading the gold and silver grades to 19.5 g/t and 430 kg/t, respectively. 96.5% of the sulphur was oxidized during the POX step.

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After the hot cure, lime boiling was conducted to break down the silver jarosite formed during POX, since it is refractory to cyanidation. The washed hot cure residue was pulped to 10% w/w solids and boiled at 95°C for 6 hours with 294 kg/t feed of CaO added, relative to the lime boil feed (46 kg/t based on the POX feed). The lime addition rate was based on the estimated sulphate concentration of the hot cure residue. Following the lime boil, the pulp was sent to CIL testing.

CIL testing of the POX-HC-LB pulp resulted in extractions of 96.9% Au and 93.4% Ag. These results compared well with the previous test program results, through which 98.2% of the contained gold and 96.6% of the contained silver were extracted (results shown in Table 13-4).

13.1.5AFRILAB 2023 Sample Campaign

The African Laboratory for Mining and Environment (AFRILAB) assayed 117 samples to determine the total sulfur content in each. Silver, zinc, lead, and iron content were determined by XRF analysis on drill cores logged by the exploration team. The results of the 2023 Sample Campaign were used to produce a sulphur content model based on elemental compositions and mineralogical associations. The multi-linear regression analysis indicated a strong correlation (R² = 0.98) between the calculated values and the actual laboratory test results, demonstrating the development of a valid relationship between the metal elements' content and the total sulfur content. The equation and graph representing this correlation are shown in Equation 13.2.1 and Figure 13-4, respectively.

Figure 13-4    Correlation Between Analyzed and Calculated Sulphur Content

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13.1.6BGRIMM 2025 Testwork Program

In 2025 BGRIMM was contracted to perform further testwork on SGS produced samples with the primary tasks included conducting fundamental test analyses of the sample and performing experimental studies on processes such as roasting and cyanidation for gold leaching of the pyrite concentrate. The objective was to determine the optimal process conditions and parameters for the main flowsheet, thereby achieving the highest leaching recovery rates for gold and silver under the best process conditions.

SGS prepared the BGRIMM sample utilising the test conditions identified in the 2022 and 2023 locked cycle tests. It should be noted the bulk flotation conditions were open circuit resulting in slightly lower gold and silver in the pyrite concentrate than would be expected.

Table 13-20    SGS Bulk Flotation Results

Product			Assays, %, g/t
			Pb			Zn			Fe			As			S=			Au			Ag
Pb Ro Conc			9.57			5.33			27.4			2.68			33.6			15.5			711
Zn Ro Conc			0.56			29.1			10.8			0.95			30			2.17			128
Py Ro Conc			0.28			0.23			40.9			1.8			44.4			4.1			46.6
Pyrite Rougher Tails			0.06			0.07			2.54			0.091			1.16			0.19			7.9
Head (calc.)			0.88			1.86			19.2			0.97			21.1			2.9			81
Head (direct)			0.88			2.45			19.7			1.03			22.5			4.11			96

13.1.6.1Feed Sample Properties

BGRIMM received 60 kg of the bulk pyrite concentrate, measuring a d(0.9) particle size of 77.53 µm, true density of 4491.2 kg/m3, and elemental analysis provided in Table 13-21.

Table 13-21    Elemental Analysis Results of Pyrite Concentrate (%)

Element			Au(g/t)			Ag(g/t)			As			Fe			S			C			Si			Al			Mg			Sb
Content			3.96			80.10			1.86			37.95			45.64			0.044			3.24			0.65			0.067			0.021
Element			Pb			Ca			Cd			Bi			Ti			Bi			Zn			Mn			Cu			Sn
Content			0.22			0.082			0.021			0.007			0.017			0.007			0.24			0.0086			0.063			0.011

X-ray diffraction analysis indicated the received sampled primarily contains pyrite as the dominant mineral phase with a minor presence of a mica mineral phase, while chemical phase analysis of gold indicated that the majority of gold is encapsulated with ~15% of the gold exposed.

13.1.6.2Roasting Condition Testing

Single Stage Roasting Condition Test

To investigate the roasting behavior of this pyrite concentrate and determine the optimal single-stage roasting temperature, single-stage roasting temperature experiments were conducted at temperatures ranging from 500 - 950℃ under a strongly oxidizing atmosphere to ensure complete reaction at each temperature.

The primary objective of roasting under a strongly oxidizing atmosphere is to remove sulfur (S). Experimental results indicate that as the roasting temperature increases, the calcine yield gradually declines, the arsenic (As) removal rate initially decreases and then rises, while the sulfur (S) removal rate generally exhibits an upward trend.

As the temperature increases, the Pb volatilization rate rises. When the roasting temperature exceeds 650℃, the Pb volatilization rate from the material accelerates significantly. Pb vapor tends to cause clogging and

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corrosion of equipment pipelines. Moreover, Pb volatilization may entrain fine gold particles into the flue gas, resulting in gold losses. Therefore, the roasting temperature should be maintained below 650℃.

Each roasting product was grinded and further subjected to 48 h leach tests at 33% w/w pulp density and 6 kg/t NaCN with the results indicated in Figure 13-5.

Figure 13-5    Au and Ag Leach Rates at Different Roasting Temperatures

As the temperature gradually increases, the Au leaching rate first rises and then declines, reaching a maximum of ~56% at approximately 630℃. In contrast, the Ag leaching rate decreases progressively with increasing temperature, peaking at ~48% at 500℃. Based on this evaluation, the optimal single stage roasting temperature for this pyrite concentrate is determined to be 630℃.

Two Stage Roasting Condition Test

Building on the single stage roasting test, a two-stage roasting temperature condition tests were conducted under varying first stage calcination temperatures (450-630 ⁰C), with the first stage using a weakly oxidizing atmosphere and the second stage a strongly oxidizing atmosphere. Each roasting step was performed for 2 hours.

The experimental results indicate that when the second-stage calcination temperature is fixed at 630℃, variations in the first-stage calcination temperature have minimal impact on calcine yield, desulfurization rate, and deleading rate. The calcine yield remains stable between 70%~71%, while the desulfurization rate consistently stays around 99.5%. However, the first-stage calcination temperature significantly affects arsenic (As) removal efficiency. The arsenic removal rate initially increases and then decreases with rising temperature, reaching a maximum of 86.34% at a calcination temperature of 580℃.

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Leaching was performed on the two stage roasted product at the same conditions used for the single stage roast products. The experimental results demonstrate that when the first stage calcination temperature is set at 530℃, the leaching rate of gold (Au) reaches its highest value of ~63%, while the silver (Ag) leaching rate under these conditions is ~46%. After comprehensive evaluation, the calcination method for this concentrate was determined to be a two-stage process at 530-630℃.

Roasting Time Condition Test

Under the optimal two stage roasting temperature conditions, the roasting time for each stage was set in 30 min increments from 30 min to 3 hours. According to the test results, as the roasting duration increases, the calcine yield gradually decreases, while the removal rates of As and S generally show an upward trend. When the roasting time exceeds 2hours, the sulfur content in the calcine essentially ceases to decrease. When the roasting time exceeds 2.5hours, the arsenic removal rate no longer shows significant improvement.

Utilizing the same leaching conditions the test results indicated, as the roasting time increases, the leaching rates of gold and silver gradually rise. When the total time of the two stage roasting exceeds 2.5 hours, the leaching rates of Au and Ag from the calcine basically stop increasing. Therefore, the optimal roasting time for each stage is determined to be 2.5 hours.

Sodium Roasting Exploratory Test

"Sodium roasting" is an important technology in mineral processing and metallurgical pretreatment. Its core idea is: at high temperatures (usually in the range of 500-1000℃), sodium salts are added to mineral raw materials, causing them to chemically react with the target components or impurities in the raw materials. This changes the chemical form or physical properties of the target components, making them easier for subsequent leaching, separation, or purification. "Sodium roasting" also plays a certain role in the oxidative pretreatment of some refractory gold ores (encapsulated in sulfides, arsenides, or silicates).

Three samplers were roasting, a no additive control sample, one with sodium carbonate, and one with sodium hydroxide at the two stage roasting conditions already determined. The results indicate that the addition of sodium salts leads to an increase in calcine yield but a decrease in the removal rates of harmful elements.

Leaching of the products indicate that soda roasting does not improve the gold leaching rate but significantly enhances the silver leaching. However, the high content of harmful elements in the iron concentrate after soda roasting compromises its quality. Consequently, the addition of sodium salts for roasting is not recommended.

Hydrated Lime Roasting Exploratory Test

Addition of 6 g of Ca(OH)2 with 100 g pyrite concentrate sample and utilizing the optimum roasting, acid-, and cyanide leaching conditions resulted in reduction of sulfur and arsenic removal rate as well as a reduction of ~12% gold extraction rate and ~14% silver extraction rate. Roasting with hydrated lime is not recommended at this stage.

13.1.6.3Cyanidation Pre-Treatment Testing

A large quantity of calcine was produced under optimal roasting conditions (530 ⁰C for 2.5 hours, and 630 ⁰C for 2.5 hours) for subsequent experiments achieving an average calcine yield rate of 69.7%.

Leach vs CIL Testing

Test conditions were grinding to P90 30µm, 33%w/w solids, pH 11, 6kg/t NaCN at 48 hour leaching, with and without 20 g/L activated carbon.

The results show either leaching method is suitable for this material, with similar extraction raters for with and without activated carbon addition.

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Grind Sensitivity Testing

The same leaching conditions were used but the grind size was varied from P90 0.3-4µm. The results indicated that the material had to be ground to below 1µm to see some improvement in gold recovery with no difference in silver recovery. The energy requirements and material handling would be cost prohibitive to explore further at industrial scale.

Cyanidation Pre-Treatment Testing

Utilizing the established leaching conditions in the leach vs CIL testing, four pre-treatments methods were tested to establish the effect on the final leach rates, namely:

1.Roasting – Cyanidation

2.Roasting – Water Quenching – Cyanidation

3.Roasting – Acid Leaching – Cyanidation

4.Roasting – Water Quenching – Acid Leaching – Cyanidation

Roasting conditions were the established optimum, and leaching conditions were pH of 1, liquid-to-solid ration of 4, at 90 ⁰C for 4 hours. Results are shown in Table 13-22.

Table 13-22    Cyanidation Pre-Treatment Evaluation

No			Calcine Yield（%）			Acid Leaching Residue Rate（%）			Gold Grade（g/t）						Gold Leaching Rate （%）			Silver Grade（g/t）						Silver Leaching Rate （%）
									Raw Material			Cyanidation Residue						Raw Material			Cyanidation Residue
1			69.73			-			3.96			1.98			65.14			80.10			58.10			49.42
2			69.73			-			3.96			1.88			66.90			80.10			59.60			48.12
3			69.73			95.34			3.96			1.82			69.45			80.10			43.00			64.31
4			69.73			94.08			3.96			1.72			71.51			80.10			43.90			64.05

Water quenching improves gold recovery by ~1% and acid leaching improves gold and silver recovery by ~5% and ~15% respectively. It is recommended to adopt the roasting-water quenching-acid leaching-cyanidation process for the pretreatment of pyrite concentrate.

13.1.6.4Acid Leaching Condition Testing

Liquid-to-Solid Ratio Testing

The liquid-to-solid ratio during acid leaching is one of the critical factors governing the effectiveness of both acid leaching and subsequent cyanidation leaching processes. If the liquid-to-solid ratio is too low, the fluidity of the acid solution becomes poor, potentially leading to insufficient contact between the acid and the slurry, incomplete reactions, and low impurity removal efficiency. Conversely, if the liquid-to-solid ratio is too high, excessive acid solution may be used, increasing the costs associated with subsequent filtration and waste acid treatment. Therefore, it is necessary to determine the optimal acid leaching liquid-to-solid ratio experimentally.

Test conditions for leaching were pH 1 with sulfuric acid, 90 ⁰C, 4 hour acid leach time with varying L/S ratios (4:1, 5:1, 6:1, and 7:1), followed by the established cyanide leaching conditions. Based on the test results a L/S ratio of 6:1 is recommended for acid leaching.

Temperature Condition Testing

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Maintaining the previous acid leaching (with L/S ratio of 6:1) and cyanidation leaching test conditions and varying the acid leach temperature from 60 – 90 ⁰C in 10 ⁰C increments, indicated that the gold and silver leach extraction gradually increase with increasing acid leaching temperature. By increase the temperature from 60 to 90 ⁰C the gold and silver leach rate increased by ~2% respectively.

Leach Time Testing

Maintaining the previous acid leaching (with 90 ⁰C Temperature) and cyanidation leaching test conditions and varying acid leach time from 1 to 5 hours in 1 hours increments, indicated that the gold and silver leach extraction increase by ~2% respectively from 1 to 3 hours with no further improvement beyond that.

13.1.6.5Cyanidation Condition Testing

A Bulk sample at the optimum roasting and acid leaching conditions were prepared.

Grind Optimization Testing

Cyanidation leach conditions were set at pH 11 with lime, 6 kg/t NaCN, 33%w/w pulp density, 48 hour leach time and varying the grind from unground (P90 80µm) to P90 38, 22, 17, and 15 µm. The gold extraction increased by ~3% between unground to P90 38µm with no further improvement in the finer grinds and the grind had no effect on the silver extraction rate.

Cyanide Dosing Optimization Testing

Cyanidation leach conditions were set at previous conditions (with P90 at 38 µm) and varying the cyanide dosage rate between 1 to 5 kg/t in 1kg/t increments. Gold extraction was found to improve by ~2% from 1 to 3 kg/t NaCN with no further improvement at higher dosages and silver had a marginal increase in extraction rate from 1 to 2 kg/t NaCN with no increase beyond that.

Leach Time Optimization Testing

Cyanidation leach conditions were set at previous conditions (with 3 kg/t NaCN dosage rate) and varying the leach time from 12 to 48 hours in 12 hour increments. The optimum leach time was determined as 36 hours with ~72% gold dissolution and ~64% silver dissolution.

13.1.6.6Verification Testing

Four tests were performed at the selected optimum conditions, and the repeatability of the results were confirmed.

From the results of the fluidized bed roasting test, in the small-scale fluidized bed roaster, the average calcine yield is 69.27%, the average removal rate of Pb is 11.83%, the average removal rate of S is 99.60%, and the average removal rate of As is 84.08%. The calcine yield and the removal rates of harmful elements are close to those of the muffle furnace roasting results.

From the cyanidation results, after the calcine roasted in the fluidized bed roaster undergoes water quenching, acid leaching, grinding and cyanidation, the average cyanide leaching rate of gold is 72.11% and that of silver is 63.48%, which are consistent with the results of muffle furnace roasting.

13.1.7Hazen 2025 Testwork Program

In 2025 Hazen Research Inc. was contracted to perform roast leach extraction testwork to recover gold and silver from pyrite flotation concentrate with a summary below extracted from the ongoing testwork campaign.

Hazen has trialed varying temperatures, roasting gas compositions, additives, and additional grinding prior to leaching. The testwork campaign is still ongoing with gold recoveries ranging from 53 to 80% and silver recoveries ranging from 25 to 86%, preliminary observations as follow:

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Reducing roast with nitrogen feed gas did not improve metal recoveries.

6% Lime addition improved silver recovery by 23.5%.

Increasing roasting temperature (600 to 800 ⁰C) reduced gold and silver recoveries by ~50% and ~40% respectively. Sintering of the material is expected to be the cause of the sharp decrease in recoveries with 600 to 650 ⁰C was chosen for follow on tests.

Sodium hydroxide shows no improvement over lime with plugging observed and agglomerates formed in the calcine product.

Crushing of the calcine (12 minutes in a ring and puck mill) prior to leaching provided inconsistent results with questionable PSD measurements of the material.

Gold and silver balances were overall poor, 77 to 125% and 69 to 111% respectively, likely due to the "nugget" effect.

13.2Results Interpretation and Conclusions

This section describes how the results presented in Section 13.2 have been interpreted and used in the development of the process design criteria. At the time that the process design criteria were being developed, only preliminary results from the SGS 2022 and SGS 2024 locked-cycle testing campaigns were available, so these results were used with the intention that future testwork results would be confirmatory. The criteria developed from these results are outlined in Table 17-1 in Chapter 17. Orway Mineral Consultants, a subsidiary of Lycopodium, provided the design for the comminution circuit.

13.2.1Mineralogy / Metallurgy

The head grades used in the process plant design were the calculated head grades from the 2022 locked cycle tests flotation tests conducted by SGS (detailed in Section 13.2.4). The selection of these grades allows for complete metallurgical accounting and direct alignment with reagents consumed during testwork and the recoveries observed.

13.2.2Comminution Design

OMC used its testwork database to infer missing data and to recommend further testwork for subsequent studies. While both BWi values from testwork fell outside the database range for other similar resources in the area, the average BWi fell at the 81st percentile of the range, aligning well with OMC’s typical strategy of designing to the 85th percentile hardness. As such, the average of both test results was used in the comminution circuit design (13.1 kWh/t). The remainder of the comminution circuit design criteria were estimated from OMC’s database.

13.2.3Flotation Circuit Design

The conditions tested during the 2022 and 2024 SGS locked cycle testing campaigns were used as the basis for the concentrate recovery circuit design, since the testing incorporated recycles at steady state and produced concentrates of acceptable grade and recovery (detailed in Section 13.1.3.5 and Section 13.1.4.3 respectively). Using locked cycle data provides a higher design certainty since it more closely represents plant operation by the incorporation of recycles and since it validates the process conditions at steady state.

The design criteria derived from the 2022 and 2024 locked cycle testing includes:

•General circuit design (i.e. orientation, number of stages, implementation of regrinding).

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•Reagent addition.

•Concentrate mass pulls, grades and recoveries.

•Residence times.

The flotation recoveries for the various metals were applied to the head grades from the mine plan to estimate annual concentrate productions. The lead, zinc and sulphur grades from the locked-cycle tests were used to determine the production rates of lead, zinc and pyrite concentrates, respectively. The gold and silver grades of each concentrate was then calculated using the recovery of the metals to each concentrate, based on the locked-cycle tests.

13.2.4Pyrite Concentrate Production

The correlation developed between the total sulphur and metals’ content in the drill core samples (Equation 13.2.1, described in Section 13.1.5) was used to develop a model to estimate the life of mine pyrite concentrate produced, assuming constant concentrate grades and recoveries based on SGS locked cycle testing (2022 and 2024). This model was applied to the project financial model to estimate pyrite concentrate production based on the mine plan. The details of the financial model are further discussed in Chapter 22.

13.2.5Pyrite Concentrate Treatment

Throughout recent testwork campaigns, several methods of oxidizing and leaching the pyrite concentrate were trialled to achieve the highest recoveries of gold and silver (while minimizing costs associated with reagent consumption, energy required, and residence time). Leaching tests to date have shown that the material is highly refractory and resistant to direct cyanidation for gold and silver recovery. A comparison of tested methods is provided in Table 13-23 (including only testwork conducted on fresh material, excluding the historic tailings).

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Table 13-23     Comparison of Pyrite Treatment Methods

Treatment Process			Testwork Campaign			Process Conditions			Sulphide Oxidation %			Au Extraction1 %			Ag Extraction1 %			NaCN Added kg/t feed			Lime Added kg/t feed			NaCN Consumed kg/t feed			Lime Consumed kg/t feed
Direct Cyanidation (CIL)			SGS 2018			30 h pre-aeration, 40% pulp density, 48-h leach			N/A			19			58			12.3			31.3			9.96			31.3
						30 h pre-aeration, 40% pulp density, 48-h leach			N/A			32			70			12.1			29.8			7.99			29.7
			SGS 2019			D80 4.47 µm, 40% pulp density			N/A			17.3			83.9			24.9			1.83			22.3			1.60
			SGS 2022			D80 42 µm, 4-h pre-aeration (pH 10.51, 20 ppm DO, 250 g/t PbNO3), 72-h leach at pH 10.5-11			N/A			9.5			36.4			20.3			4.25			4.39			3.91
						D80 6.3 µm, 4-h pre-aeration (pH 10.51, 20 ppm DO, 250 g/t PbNO3), 72-h leach at pH 10.5-11			N/A			22.4			80.2			28.5			15.4			14.5			15.3
			Mintek 2024			24-h leach, 25% solids, 20 g/L CN, DO>10.5, then 8-h CIL			N/A			29			40			-			-			2.11			39.1
POX-CIL			SGS 2018			5.6% pulp density, pre-acidulation (pH 1.25), 225°C, 690 kPa, 120 min			99			97.4			49.2			7.24			320.8			3.55			320.8
						5.6% pulp density, pre-acidulation (pH 1.39), 225°C, 690 kPa, 60 min, hot cure			99			97.6			1.2			6.56			16.8			4.45			16.8
						5.6% pulp density, pre-acidulation (pH 1.30), 225°C, 690 kPa, 60 min, hot cure and lime boil			-			98.2			97.6			2.11			0.6 + 59.2 kg/t for lime boil			0.90			0.6 + 59.2 kg/t for lime boil
			SGS 2019			5.5% pulp density, 30-min pre-acidulation (pH 1.50), 225°C, 690 kPa, 60 min, hot cure and lime boil			99.8			98.2			96.6			2.23			3.98			1.52			3.98 + 35.8 kg/t for lime boil
			SGS 2024			8% pulp density, 30-min pre-acidulation (pH 1.50), 225°C, 690 kPa, 60 min, hot cure and lime boil			96.5			96.9			93.4			2.52			0 + 46.1 for lime boil			0.58			0 + 46.1 for lime boil
Albion-CIL			SGS 2019			52-h, 95°C, 10% pulp density, pre-acidulation (pH 4.5), O2 sparge 1 L/min, limestone addition to maintain pH 5.5, 23 kg/t H2SO4 and 440 kg/t CaCO3			28%			45.4			84.8			3.85			11.7			3.52			11.6
						72-h, 95°C, 10% pulp density, pre-acidulation (pH 4.5), O2 sparge 1.7 L/min, limestone addition to maintain pH 5.5, 21 kg/t H2SO4 and 1,209 kg/t CaCO3			68%			81.0			92.3			5.46			6.79			5.11			6.79
			SGS 2022			96-h, 95°C, D80 6.3 µm, 10% pulp density, limestone addition to maintain pH 5.5			96.3			94.4			94.4			7.92			2.48			3.22			2.40
						72-h, 95°C, D80 10 µm, 10% pulp density, limestone addition to maintain pH 5.5			74.7			84.9			91.3			8.25			3.32			2.29			3.27

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Treatment Process			Testwork Campaign			Process Conditions			Sulphide Oxidation %			Au Extraction1 %			Ag Extraction1 %			NaCN Added kg/t feed			Lime Added kg/t feed			NaCN Consumed kg/t feed			Lime Consumed kg/t feed
Roast-CIL			SGS 2018			Two-stage roast, 2 hours per roast, at 400°C and 700°C			49.6			62.5			9.1			2.82			1.16			1.02			1.16
						Two-stage roast, 2 hours per roast, at 400°C and 700°C, ground to <10 µm before CIL			-			84.8			72.5			4.17			4.29			3.91			4.27
			SGS 2022			Two-stage roast, 2 hours per roast, at 400°C and 700°C			99.9			61.1			19.9			21.9			12.0			6.54			12.0
						Two-stage roast, 2 hours per roast, at 400°C and 700°C. Pre-CIL re-grind to 7.4 µm			99.9			71.1			44.9			24.0			6.88			4.83			6.63
						Two-stage roast, 2 hours per roast, at 400°C and 600°C			97.6			55.6			51.3			24.1			N/A			13.3			N/A
						Two-stage roast, 2 hours per roast, at 400°C and 600°C. Pre-CIL re-grind to 6.7 µm			97.6			65.9			60.3			31.4			12.2			12.7			12.2
						Three-stage roast, 2 hours per roast, at 400°C and 600°C (stage 2 and 3). Aeration during 2nd and 3rd stage (1 L/min and 2 L/min)			99.8			59.3			32.5			27.3			8.56			10.3			8.56
						Three-stage roast, 2 hours per roast, at 400°C and 600°C (stage 2 and 3). Aeration during 2nd and 3rd stage (1 L/min and 2 L/min). Pre-CIL re-grind to 7.5 µm			99.8			66.7			51.6			27.3			5.93			5.10			5.54
						One-stage roast, 2 hours, at 500°C			89.3			52.5			51.8			29.1			34.7			13.6			34.7
						One-stage roast, 2 hours, at 500°C. Pre-CIL re-grind to 8.1 µm			89.3			63.9			61.8			37.2			14.6			22.5			14.6
						One-stage roast, 2 hours, at 600°C			95.1			56.4			29.6			30.5			13.5			13.9			13.5
						One-stage roast, 2 hours, at 600°C. Pre-CIL re-grind to 5.9 µm			95.1			66.7			50.4			26.8			9.3			11.8			9.1
			SGS 2024			Two-stage roast, 2 hours per roast, at 400°C and 700°C			99.9			73.0			11.3			5.84			5.97			0.62			5.97
						Two-stage roast, 2 hours per roast, at 400°C and 700°C. Pre-CIL re-grind to 4 µm			99.9			80.4			30.2			7.6			5.39			2.81			5.39
			BGRIMM 2025			Roasting: Mild oxidative roasting (530 ⁰C, 2.5-h), Intensive oxidative roasting (630 ⁰C, 2.5-h); Water Quenching; Acid Leaching: pH 1, 90 ⁰C, L/S 6:1, 3-h; Cyanide Leaching: P90 38 µm, 3 kg/t NaCN, 36-h, 33% pulp density			99.6			72.0			62.9			3			N/A			N/A			8.75 g/L (Acid Leach) 3.28 kg/t (NaCN Leach)
			Hazen 2025			Best 3 Tests: Roast at 600 ⁰C with 6% Lime and Air with 10%SO2 gas, crushed, leached, recrushed and releached.			N/A			77.5			84.0			N/A			N/A			25.7			9.1
						Average Tests: Roast at 600 ⁰C with 6% Lime and Air with 10%SO2 gas, crushed, leached, (excluding: recrushed and releached)			N/A			54.4			71.8			N/A			N/A			8.4			14.1

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Treatment Process			Testwork Campaign			Process Conditions			Sulphide Oxidation %			Au Extraction1 %			Ag Extraction1 %			NaCN Added kg/t feed			Lime Added kg/t feed			NaCN Consumed kg/t feed			Lime Consumed kg/t feed
BIOX-CIL			SGS 2022			24-h stabilization, 18-day residence, 10.5% pulp density, 40°C, air sparge, pH 1.6			60.6			75.2			75.5			17.7			9.01			7.91			8.77
						24-h stabilization, 26-day residence, 10.5% pulp density, 40°C, air sparge, pH 1.6			88.0			82.2			40.1			15.3			11.4			8.75			11.2
						24-h stabilization, 26-day residence, 10.5% pulp density, 40°C, air sparge, pH 1.6			92.8			85.1			37.0			23.6			16.2			23.4			16.2
MACH Reactor-CIL			Mintek 2024			6-h pre-oxidation, 4-h boosted leach, 72-h bottle roll cyanidation			16-18			40			70			-			-			36			273
						22-h pre-oxidation, 4-h boosted leach, 72-h bottle roll cyanidation			54			60			75			-			-			32			639

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13.3Future Testwork Recommendations

The following testwork is recommended to progress the design to the next phase:

Sample Requirement:

•Spatial composites to represent different physical locations throughout the deposit

•Samples from different zones and depths for variability testing.

•Sufficient sample sizes to conduct downstream testwork.

•Number of representative samples to be determined by AGS.

•Sufficient sample size to generate representative concentrate samples for off-taker evaluation.

Comminution Testwork:

•Additional tests to determine:

•Crushing work index (CWi)

•SMC parameters: abrasivity (Ai) and impact breakage (A x b)

•BWi (confirmatory)

•Variability of these parameters throughout the deposit and determine correlations to lithology.

•Flotation regrind tests to determine specific energy; engage with vendors for criteria relating to equipment sizing and specifications.

Ore Characterization Testwork:

•Assess material variability by assaying samples from different zones without blending according to mine plan and to create design ranges.

•Particle size distribution of ROM material and of final concentrates in support of dewatering testwork.

•Rheology testwork to determine apparent viscosities at varying shear rates and varying slurry % solids in support of slurry pipe and pump design.

•Physical characteristics impacting material handling, such as feed moisture content, specific gravity, bulk density, and stacking and clumping / compaction properties.

Flotation Testwork:

•Variability testing – open circuit and locked cycle to determine the impacts of feed variance and process upsets on grades and recoveries. Develop relations between open circuit and locked cycle results

•Produce enough of each concentrate to support dewatering testwork and further pyrite concentrate treatment testwork.

Dewatering Testwork:

•Particle settling testwork to determine the flux rate for thickener sizing.

•Flocculant screening tests.

•Filtration testwork to determine:

•Filter type

•Optimum feed pulp density

•Achievable cake moisture levels

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•Transport moisture limit targets

•Filter cycle times

•If any filter aids are required.

Pyrite Concentrate Treatment Testwork:

•Continue development of pyrite oxidation option.

•Assess variability of pyrite concentrate via spatial composites and variability testing.

•Generate sufficient pyrite concentrate for off-taker evaluation.

Water Analysis:

•Surface water quality analysis.

•Ground water quality analysis.

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14.0MINERAL RESOURCE ESTIMATES

14.1Introduction

This section describes the 2026 MRE methodology and summarises the key assumptions adopted by SRK during preparation. The resource database provided by Aya to support the MRE was independently reviewed and verified by SRK. SRK considers that the 2026 MRE reported herein provides a reasonable representation of the in situ Ag, Au, Cu, Zn and Pb mineralization based on the current level of geological understanding and the available drilling, sampling and analytical data.

SRK also estimated arsenic (As) and iron (Fe) for mine planning and technical study purposes. However, owing to the lower level of data support for As and the potential for these elements to be associated with multiple mineral assemblages (specifically Fe), SRK has not assigned a Mineral Resource classification to the As and Fe estimates.

A total of 940 drillholes, comprising 356,323.1 m of drilling, are located within the Boumadine Mining Licence. Of these, 855 drillholes, totalling 319,464.8 m, were completed to support Mineral Resource definition and extensions to the known mineralization, which is shown in Figure 14‑1, while the remaining 85 drillholes relate to prospecting within the Mining Licence and did not specifically target the known Boumadine mineralization.

The MRE has been prepared by Guy Dishaw (Principal Consultant, SRK) and James Williams (Senior Resource Geologist) in accordance with the CIM Definition Standards for Mineral Resources and Mineral Reserves (May 2014) and National Instrument 43-101 – Standards of Disclosure for Mineral Projects (May 2016). Guy Dishaw, P.Geo., is a Qualified Person ("QP") for Mineral Resource reporting and completed a site visit to the Project accordance with the requirements of NI 43-101 (Section 12.2.1).

The MRE has been prepared using satellite imagery, core photographs, geological mapping, DD, MP drilling, RC drilling, structural measurements, weathering interpretations, historical as-built depletion solids and associated mine plans, Aya's previous mineralization models, and other supporting geological datasets available.

Leapfrog Geo was used to construct the geological and mineralization models (estimation domains) and prepare the assay data for geostatistical analysis. Maptek Vulcan (Vulcan) was used to construct the block model, estimate metal grades and density values, and incorporate historical underground depletion. Supervisor software was used to assess spatial grade continuity through geostatistical analysis, whilst a combination of Leapfrog, Datamine and Vulcan was used to independently validate the MRE where appropriate. Vulcan was subsequently used to report the Mineral Resource Statement presented in Section 14.2.

Aya developed a Net Smelter Return (NSR) and mining cut-off value calculation for both open pit and underground reporting for the Project, which was independently reviewed by SRK and adopted for reporting the 2026 MRE (Section 14.15). The NSR calculation was used as the basis for the conceptualized open pit optimization, which was undertaken by Aya using Whittle. The open pit optimization was reviewed by SRK. In addition, the NSR values were used to generate underground reporting shells defining potentially mineable mineralization by underground mining methods (small isolated volumes removed).

14.2Mineral Resource Estimation Procedure

The MRE methodology adopted for the Boumadine deposit broadly comprised the following steps:

•compilation, review and validation of the geological, drilling, sampling and analytical databases;

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•review of geological mapping, satellite/hyperspectral imagery, geophysical dataset, structural mapping and oriented core data to characterise the geological controls on the orientation, continuity and geometry of the mineralization;

•construction of weathering and oxidation (redox) model;

•construction of geological and estimation domain wireframes (mineralization wireframes), guided by geological mapping, structural measurements from oriented drill core, historical mine plans and as-built depletion surveys, together with Aya’s previous geological interpretations;

•conditioning of the assay database, compositing, metal accumulation (grade × thickness) calculations and assessment of high-grade outliers;

•geostatistical analysis and variogram modelling using the metal accumulation data;

•construction of three-dimensional block models using parent block dimensions of 5 m × 15 m × 15 m within the mineralized zones and 90 m × 90 m × 90 m for the surrounding waste;

•interpolation of Ag, Au, Cu, Pb, Zn and true thickness (TT) using Ordinary Kriging (OK), whilst Fe and As were estimated using Inverse Distance Weighting (IDW2). All analytes were estimated using the metal accumulation other than TT;

•estimation of density using single drillhole composites representing the across-strike mineralized intersection, with factoring applied within the transitional weathering domain. Density values were based on measured density data supplemented by density values derived from regression analysis undertaken by SRK;

•visual and statistical validation of the block model against the input composite data, including comparison of estimated metal accumulation values, density composites and back-calculated grades against the corresponding input data

•classification of the block model into the Indicated and Inferred Mineral Resource categories based on data quality, geological continuity and drillhole spacing;

•assessment of the reasonable prospects for eventual economic extraction (RPEEE) through the determination of appropriate reporting cut-off values and the generation of a conceptualised optimized open pit shells and underground reporting shapes;

•reporting of the Mineral Resources in accordance with the definitions, terminology and guidelines of the CIM Definition Standards for Mineral Resources and Mineral Reserves (May 2014);

•comparison of the current Mineral Resource Statement (2026 MRS) with the previous MRS prepared by Aya.

14.3Database Construction, Validation and Comments

Electronic data files were provided to SRK by Aya in the form of ASCII exports and a master Microsoft Excel workbook generated from the Company's Geotic drillhole database on 16 April 2026. The database was last updated with assays issued from Afrilabs on the 28 February 2026. The drillhole data were imported into Leapfrog Geo, where a series of validation checks was undertaken to identify potential errors, including duplicate records, overlapping intervals, missing data, invalid coordinates and other database inconsistencies.

The database supplied by Aya comprised the following information:

•drillhole collar information, including coordinates, elevation, total depth and drilling dates;

•downhole survey data;

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•lithological logging;

•geotechnical logging, including core recovery and Rock Quality Designation (RQD);

•weathering logging;

•alteration logging;

•mineralization logging;

•structural measurements collected from oriented drill core;

•density measurements; and

•sample intervals and analytical results for all recorded analytes, including Ag, Au, As, Cu, Fe, Pb and Zn.

Aya informed SRK that, where reassaying and relogging were completed for the post-2018 drillholes, the reassay and relogging data were used in place of the original assay and logging data within the Project database.

SRK did not undertake a complete audit of the original exploration records against the electronic database. However, as discussed in Section 10, SRK independently verified selected drillhole collars during the site visit, reviewed representative drill core, geological logging, sampling procedures, orientated core measurements as well as completed high level geological mapping of the mineralization exposed at surface and in the artisanal working. In addition to this SRK also checked a selection of 2025 and 2026 Afrilabs assay certificates against the database provided by Aya. SRK considers the database to be of sufficient quality to support the 2026 MRE.

A total of 852 DD, 33 RC and 55 MP drillholes, comprising 328,719.4 m, 7,513.1 m and 20,090.6 m of drilling, respectively, were available for the 2026 MRE (Table 14‑1 and Figure 14‑1), these drillholes were all located within the Boumadine Mining Licence. The database contains 329,586 Ag (g/t), 329,586 Au (g/t), 329,586 Cu (%), 329,586 Pb (%) and 329,586 Zn (%) assay records.

The historical drillholes, 2017 Maya drillholes and specific tailings-related drilling were excluded from this total. The exclusion of this data is consistent with the methodology adopted in previous MRE completed for the Boumadine Project.

Table 14‑1    Summary of the Aya drillhole database within the Boumadine Mining Licence

Year			Number of drillholes									Total number of all drillholes types			Total meters drilled									Total meters drilled (all drillhole types)
			DD			MP (RC/DD)			RC						DD			MP (RC/DD)			RC
2018			42									42			4,591.5									4,591.5
2019			24									24			3,610.8									3,610.8
2020			6									6			714.1									714.1
2022			94									94			21,543.5									21,543.5
2023			197									197			75,450									75,450
2024			191			14			3			208			91,420.8			6,202.1			469			98,091.9
2025			192			41			30			263			92,709.8			13,888.5			7,044.1			113,642.4
2026			106									106			38,679									38,679
Grand Total			852			55			33			940			328,719.4			20,090.6			7,513.1			356,323.1

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Figure 14‑1    Plan view of drillhole traces (coloured grey) and collars used in the 2026 MRE( coloured by Year) in relation to the Boumadine Mining Licence (black) and resource definition drilling area (red)

14.4Exploratory Data Analysis

SRK reviewed the distributions of Ag, Au, Cu, Pb and Zn grades relative to the logged lithology and observed that elevated metal grades are predominantly associated with the massive sulphide (MS; >70% logged visible sulphide content), semi-massive sulphide (SMS; <70% logged ~>30% sulphide content) and gossan (A3)

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lithologies (Figure 14‑2). However, significant mineralization also occurs within other logged lithologies, reflecting the complex distribution of sulphide mineralization and local inconsistencies in the geological logging.

To provide a consistent guide for mineralization wireframing, SRK generated a calculated total sulphide variable for each assay interval. The calculation was used only as a geological modelling aid where mineralized intervals had not been logged as MS, SMS or A3, and was not used as a substitute for geological interpretation, it was a guide to be used alongside the assay intervals.

The total sulphide variable was derived using stoichiometric relationships based on the principal sulphide minerals identified during previous mineralogical investigations at Boumadine. Copper, zinc and lead were assumed to occur predominantly as chalcopyrite (CuFeS₂), sphalerite (ZnS) and galena (PbS), respectively. The proportion of pyrite (FeS₂) was subsequently estimated from the remaining Fe after deducting the iron attributed to chalcopyrite and sphalerite. One of the key assumptions is that Fe was not associated with any silicate minerals (primary or alteration) in a large abundance. The individual calculated sulphide percentages were then summed to produce a total sulphide estimate for each sample. The sulphide calculations are noted below.

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Figure 14‑2    Box plot showing raw Ag, Au and Zn grades with Aya lithological logging codes (A= Ag g/t, B= Au g/t and C = Zn %)

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14.5Topography and Collars Adjustments

The geological, weathering and mineralization domain models described in the following sections were constructed beneath a topographic surface wireframe (Figure 14‑3) that exhibits moderate local relief. The northern part of the deposit is characterised by a series of north–south-trending ridges, which have similar trends to the mineralized structures, whereas the central and southern portions of the deposit occur within a broader river valley and generally display more subdued topography. In the far southern part of the Mining Licence, the relief becomes more pronounced, similar to that in the north.

Aya acquired the topographic surface from an open-source satellite-based digital elevation model (DEM) with a reported spatial resolution of 5 m, in addition to this Aya applied a vertical correction of 59.3 m.

SRK visually reviewed the surface drillhole collar positions relative to the topographic surface and noted an overall reasonable correlation between the surveyed collar elevations and the digital terrain model (DTM). However, several localized discrepancies were identified, with some drillhole collars differing from the topographic surface. SRK considers these discrepancies typically reflect limitations of the existing topographic dataset rather than inaccuracies in the surveyed collar positions.

To provide a consistent basis for geological modelling, surface drillholes without surveyed collar coordinates (i.e. where both the Survey and Survey Type fields were blank or noted as planned in the database) were projected vertically onto the topographic surface. Surface drillholes with surveyed collar coordinates were retained at their recorded locations, with the exception of BOU-DD24-455, which is a daughter hole, the underground drillholes and channel samples were retained at their original surveyed coordinates. All historical and 2017 surface drillholes were projected onto the topographic surface, irrespective of their recorded collar elevations, owing to the lower confidence associated with the historical topographic data.

SRK considers the resulting topographic surface and to be sufficiently accurate for the purposes of the 2026 MRE and considers the available collar survey data to be appropriate for the current level of study. However, Aya plan to undertake a high-resolution pre-mining drone survey, once this has been acquired this should be used in future MRE to improve the accuracy of the surface topography for subsequent FS work.

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Figure 14‑3    Boumadine topographic wireframe in plan view coloured by elevation and drillhole collars (coloured by year)

14.6Geological Modelling and Domaining

The Boumadine geological modelling is primarily based on the Aya drilling database, and is also supported by historical underground mine plans (includes geological mapping information), surficial geological mapping, as-built depletion surveys, satellite imagery, structural measurements and Aya's previous mineralization models. The maps (surface and underground) have been georeferenced and have been integrated into the lithological and mineralization models.

Boumadine is interpreted as being formed within a shallow submarine volcanic environment developed in an extensional basin or graben setting. Mineralization generally comprises 1 to 5 m-wide (although locally thicker) MS and or SMS veins and lenses that dip approximately 70° towards the northeast, with dip azimuths ranging from 050° to 060°. The MS and SMS veins is generally surrounded by a 1 to 10 m (locally up to 20 m) thick halo of 10 to 30% disseminated pyrite.

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14.7Lithological Model

No three-dimensional lithological models were constructed by SRK for the 2026 MRE. This also includes the overburden, which is interpreted to be restricted to the valley floors, where it has limited lateral extent and thicknesses.

As noted in Section 7.2, the distinction between the felsic and mafic tuffs is an important component of the geological interpretation and appears to influence the distribution and intensity of mineralization; however, this distinction is locally difficult to recognize or is absent from the historical drillhole logging and therefore based on this SRK recommends that Aya should investigate whether whole rock geochemistry will help with the identification and standardisation of the lithologies encountered in the drilling.

For future MRE updates, SRK recommends that a three-dimensional (3D) lithological model is developed, subject to the availability of sufficient and consistent geological data. Such a model would improve the geological understanding of the deposit, support mine planning and geotechnical assessments, and provide a stronger geological framework for future MRE.

14.8Weathering Model

SRK reviewed the weathering logging during the site visit and subsequently undertook a desktop review of this data using the 2024 to 2026 drill core photographs in Imago. Based on this review, SRK considers the interpreted base of oxidation (top of fresh rock, coded as SU in the Aya weathering logging scheme) to be reasonably reliable, with the majority of weathered intervals are logged as TR (transitional). However, the distinction between oxide and transitional material could not be reliably established because only a limited number of intervals were logged as fully oxidised (OX), and in several cases these occurred at depth rather than within the near-surface weathering profile, suggesting that they may instead represent localized fault-related oxidation. Furthermore, on several occasions the logged TR and SU contacts did not correspond with SRK's visual assessment of the drill core or surrounding drillholes. In particular, SRK identified several instances where overburden had been incorrectly logged as weathered bedrock, or where TR was logged for entire drillhole or large lengths of core, though the adjacent holes only had a few meters of TR logged. These observations were communicated to Aya during the QP site visit, and Aya agreed to review and update the affected intervals within the geological database and implement consistent weathering logging procedures for future drilling programs.

SRK was only able to model two weathering domains: oxide (logged as TR in the drillhole file) and fresh (logged as SU in the drillhole file). Where discrepancies were identified, the weathering interpretation was guided by direct observations of the drill core photographs in conjunction with the surrounding drillholes.

The weathering surfaces were constructed in Leapfrog Geo based on the interpreted weathering logging (Figure 14‑4). The base of oxidation was modelled across the entire area (not split by the three areas) as a deposit offset surface constrained, with the topographic surface used as a reference. SRK considers the resulting surface to provide a reasonable representation of the oxidation profile for the purposes of the 2026 MRE.

SRK recommends that future surface drilling programs include some total sulphur (S%) analysis and, where practicable, this would also help expand the current S% database which contains 92 values currently. These data would provide an independent means of validating and refining the interpreted oxidation profile and improve the definition of the transition between oxide and fresh sulphide mineralization in the locations where this data is recorded.

Only a few mineralized samples occur within the weathered zone compared with the fresh sulphide domain, making it difficult to assess the extent to which weathering has modified the primary metal distribution other than assessing surface grab samples (not used in this MRE). Previous reports indicate that supergene

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weathering has preferentially leached mobile elements such as Zn, Mn, Ni, Co and Cd, whereas Ag, Au and Pb are interpreted to have remained relatively immobile. Consequently, the metal distribution within the weathered profile may differ locally from that observed in the underlying fresh sulphide mineralization, though further work is required to resolve this aspect.

Figure 14‑4    Cross section (30m slice) through weathering model, with drillholes coloured by logged weathering codes (TR = blue, and SU = green) and drillhole collars coloured by year (section X= 317165 and Y = 3474300)

14.9Mineralization Model

SRK modelled the mineralization for the three areas (north, central and south - Figure 14‑5.) of the deposit. Modelling the mineralization by area allowed the interpretation to better reflect local geological controls and provided greater flexibility during geological modelling and grade estimation.

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The mineralization domains comprise discrete massive to semi-massive sulphide veins. SRK did not model a separate lower-grade disseminated mineralization halo. Where disseminated mineralization exceeded the modelling cut-off value and demonstrated sufficient geological continuity, these intervals were incorporated into the domain wireframes.

As noted above, the orientations and geometries of the individual mineralization wireframes were guided by a combination of satellite imagery, surface and underground geological mapping, historical as-built depletion surveys, drillhole geology, structural logging data, logged mineralization style (including massive and semi-massive sulphides), assay data, Aya previous wireframes and SRK’s interpreted total sulphide content. Mineralization wireframes were typically modelled where mineralized intervals could be correlated across multiple drillhole sections. An iterative modelling approach was adopted; whereby massive to semi-massive mineralized cores were initially interpreted before progressively reducing the modelling thresholds to capture the broader extent of the mineralized structures. SRK also considered the NSR thresholds used for Mineral Resource reporting (previous and current) during the modelling process to ensure that potentially economically relevant mineralization was appropriately represented. These thresholds were used as modelling guides rather than strict cut-offs, allowing the interpreted wireframes to maintain geological continuity through locally lower-grade intervals were supported by the geological interpretation. In most cases, the interpreted vein contacts correspond to relatively sharp changes in grade and sulphide abundance; however, locally more gradational contacts were observed, typically reflecting a progressive decrease in sulphide abundance away from the massive to semi-massive sulphide mineralization.

SRK notes that drilling locally intersected numerous thin, discrete mineralized structures that could not be confidently correlated between drillholes or modelled owing to insufficient geological support. The geological interpretation is also subject to several limitations, including uncertainty in correlating isolated high-grade intersections; the absence of structural measurements from oriented drill core in some areas; uncertainty associated with some historical drillhole surveys; the positional accuracy of historical underground mapping; locally wide drillhole spacing relative to the dimensions of the mineralized structures; and the effects of faulting on mineralization continuity. Where supported by the available geological evidence, SRK considered the interpreted faulting when defining the extents and continuity of the mineralization wireframes. Additional infill drilling is recommended to improve geological confidence and further constrain and refine the wireframe interpretations in these areas.

The interpreted mineralization wireframes were reviewed with Aya, who provided geological feedback based on their knowledge of the deposit. This review resulted in several minor refinements to the wireframe interpretations, including the addition of a small number of discrete mineralized wireframes. In total, SRK modelled 158 mineralized domains across the Project (Figure 14‑5 and Figure 14‑6), comprising 57 domains in the southern area, 58 in the central area and 43 in the northern area.

As part of the model validation process, SRK compared the orientations of the interpreted mineralization wireframes against available structural measurements and geological mapping to assess whether the modelled geometries were consistent with the observed structural framework (Figure 14‑7). Only structural measurements assigned confidence ratings of 2, 3 and 4 are presented. SRK notes that some of the structural measurements represent subsidiary veinlets associated with the broader mineralized sulphide structures rather than the principal mineralized structures themselves. As noted by Gaboury, these subsidiary veinlets can exhibit variable orientations. Consequently, some variation between the measured structural orientations and the orientations of the modelled mineralization wireframes is expected.

The mineralization remains open at depth and along strike in several areas. To the north, further extension of the interpreted mineralization is constrained by the northern boundary of the Mining Licence. Drilling completed to date has identified mineralization over an approximate strike length of 5.8 km and, locally, to depths exceeding 1,000 m below surface, although mineralization does not extend to these depths throughout the deposit.

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Figure 14‑5    Plan view showing location of the North, Central and Suther Wireframes as modelled by SRK in relation to the Boumadine Mining Licence

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Figure 14‑6    Boumadine modelled wireframes in (light brown) in relation to the drillhole traces (grey)

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Figure 14‑7    Comparison of MS and SMS structural measurements (A, C and E) with the average orientations of the modelled vein wireframes (B, D and F) for the northern (A and B), central (C and D) and southern (E and F) areas of the Boumadine deposit.

14.9.1Estimation domains

To support grade interpolation, the mineralization wireframes were grouped into estimation domains based on their natural structural associations (dip and dip direction) within the northern, central and southern areas of the Boumadine mineralization framework. The northern area was subdivided into seven groups (N1 to N7), the central area into eight groups (C1 to C8), and the southern area into nine groups (S1 to S9). The resulting estimation domains and their respective group assignments are listed in Table 14‑2, and are illustrated on stereonets in Figure 14‑7.

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Table 14‑2    Final estimation domains

Area

Group

Wireframe Name

Estimation Domain Code

Area

Group

Wireframe Name

Estimation Domain Code

Area

Group

Wireframe Name

Estimation Domain Code

Central

C1

C_Main

2091

Central

C5

C1_West2

2502

South

S1

S_East_1a

1031

Central

C1

C1_East12

2012

Central

C5

C1_West38

2538

South

S1

S_East_2

1040

Central

C1

C1_East18

2018

Central

C5

C1_West7

2507

South

S1

S_East_2a

1041

Central

C1

C1_East19

2019

Central

C6

C1_East14

2614

South

S1

S_East_3

1003

Central

C1

C1_East2

2002

Central

C7

C1_West29

2729

South

S1

S_East_3a

1033

Central

C1

C1_East20

2020

Central

C8

C_NS2

2802

South

S1

S_East_4

1004

Central

C1

C1_East26

2026

North

N1

N_NE_1

3001

South

S1

S_East_5

1005

Central

C1

C1_East4

2004

North

N1

N_NE_10

3010

South

S1

S_East_6

1006

Central

C1

C1_East5

2005

North

N1

N_NE_11

3011

South

S1

S_East_7

1007

Central

C1

C1_East6

2006

North

N1

N_NE_2

3002

South

S1

S_Main

1090

Central

C1

C1_East9

2009

North

N1

N_NE_3

3003

South

S1

S_Main_a

1091

Central

C1

C1_West1

2101

North

N1

N_NE_4

3004

South

S1

S_Main_b

1092

Central

C1

C1_West13

2113

North

N1

N_NE_5

3005

South

S1

S_Target_1

1061

Central

C1

C1_West15

2115

North

N1

N_NE_7

3007

South

S1

S_Target_2

1062

Central

C1

C1_West17

2117

North

N1

N_NE_8

3008

South

S1

S_Target_3

1063

Central

C1

C1_West18

2118

North

N1

N_NE_9

3009

South

S1

S_West_1

1050

Central

C1

C1_West20

2120

North

N2

3_Imarirren

3191

South

S1

S_West_1a

1051

Central

C1

C1_West22

2122

North

N2

3_Imarirren East

3192

South

S1

S_West_1b

1052

Central

C1

C1_West27

2127

North

N2

3_Imarirren West

3193

South

S1

S_West_1d

1053

Central

C1

C1_West3

2103

North

N2

3_NS 0

3100

South

S1

S_West_2

1054

Central

C1

C1_West30

2130

North

N2

3_NS 1

3101

South

S1

S_West_2a

1055

Central

C1

C1_West31

2131

North

N2

3_NS 2

3102

South

S1

S_West_2b

1056

Central

C1

C1_West32

2132

North

N2

3_NS 21

3121

South

S2

S_N30_1

1201

Central

C1

C1_West4

2104

North

N2

3_NS 22

3222

South

S2

S_N30_2

1202

Central

C1

C1_West6

2106

North

N2

3_NS 23

3223

South

S2

S_N30_3

1203

Central

C2

C1_East13

2213

North

N2

3_NS 24

3224

South

S2

S_N30_8

1208

Central

C2

C1_East15

2215

North

N2

3_NS 25

3225

South

S2

S_N30_9

1209

Central

C2

C1_East16

2216

North

N2

3_NS 26

3226

South

S3

S_East_1d

1331

Central

C2

C1_East22

2222

North

N2

3_NS 3

3103

South

S3

S_East_2d

1341

Central

C2

C1_East23

2223

North

N2

3_NS 4

3204

South

S3

S_N30_10b

1310

Central

C2

C1_East3

2203

North

N2

3_NS 5

3205

South

S3

S_N30_11b

1311

Central

C2

C1_East7

2207

North

N2

3_NS 6

3206

South

S3

S_N30_12b

1312

Central

C2

C1_West19

2219

North

N2

3_NS 7

3207

South

S3

S_N30_13b

1313

Central

C2

C1_West21

2221

North

N2

3_NS_27

3227

South

S3

S_N30_14b

1314

Central

C2

C1_West24

2224

North

N2

3_NS_30

3230

South

S3

S_N30_4d

1304

Central

C2

C1_West25

2225

North

N2

3_Tizi West

3292

South

S3

S_N30_5b

1305

Central

C2

C1_West35

2235

North

N2

3_Tizi West 01

3293

South

S3

S_N30_6b

1306

Central

C2

C1_West36

2236

North

N2

3_Tizi_West_3

3294

South

S4

S_East_3d

1403

Central

C3

C1_East11

2311

North

N2

4_Tizi Main

3291

South

S4

S_East_4b

1404

Central

C3

C1_East27

2327

North

N3

3_Tizi East

3301

South

S4

S_East_7b

1407

Central

C3

C1_West_14

2314

North

N3

3_Tizi East 0

3295

South

S4

S_Main_d

1094

Central

C3

C1_West9

2309

North

N3

N_NE_6

3006

South

S5

S_SW_1

1501

Central

C4

C_NS1

2400

North

N4

1_West2

3402

South

S5

S_SW_2

1502

Central

C4

C1_East1

2401

North

N4

3_NS 10

3410

South

S5

S_SW_3

1503

Central

C4

C1_East10

2410

North

N4

3_NS 9

3409

South

S5

S_SW_4

1504

Central

C4

C1_East17

2417

North

N4

3_NS_11

3411

South

S6

S_East_14

1614

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Area

Group

Wireframe Name

Estimation Domain Code

Area

Group

Wireframe Name

Estimation Domain Code

Area

Group

Wireframe Name

Estimation Domain Code

Central

C4

C1_East24

2424

North

N5

3_NS 8

3508

South

S6

S_East_1b

1631

Central

C4

C1_East25

2425

North

N6

1_Main

3600

South

S6

S_East_1c

1632

Central

C4

C1_West11

2411

North

N7

4_EW_1

4003

South

S6

S_Main_c

1093

Central

C4

C1_West5

2405

South

S1

S_East_1

1030

South

S6

S_West_1c

1601

Central

C4

C1_West8

2408

South

S1

S_East_11

1011

South

S7

S_EW_2

4002

Central

C5

C1_West10

2510

South

S1

S_East_12

1012

South

S8

S_SW_5

1505

South

S1

S_East_13

1013

South

S9

S_EW_1

4001

14.10Data Conditioning and Post-domaining Statistical Analysis

14.10.1Grade estimation methodology

The interpreted mineralization at Boumadine comprises a series of narrow vein-like bodies displaying strike continuity significantly greater than their true thickness. Where mineralized domains are substantially thinner than the parent block dimensions, without careful selection of estimation parameters conventional 3D grade estimation can result in irregular block grade distributions, localized bias, and may give the impression of greater selectivity and predictability of grade in the across-strike dimension than is realistic.

For this reason, SRK adopted a two-dimensional estimation methodology, which estimates true thickness (TT) and metal accumulation independently before calculating block grades. SRK considers this approach to be more appropriate for the style of mineralization encountered at Boumadine, as it simplifies estimation decisions, reduces artefacts associated with estimating thin volumes in 3D, and allows for greater consideration towards accurately defining the distribution of mineralization in the plane of greatest continuity.

SRK used statistics to confirm the suitability of the domain models for use in grade estimation. Assayed Au, Ag, Cu, Pb and Zn grades were considered in this analysis.

14.10.2Data conditioning and top cutting of outliers

Prior to geostatistical analysis, SRK reviewed the grade distributions to determine whether top-cutting was required to reduce the influence of isolated high-grade outliers on the estimation. This was undertaken on the raw assay data, prior to calculation of metal accumulation and TT values.

As over 70% of the raw assay samples in the drillhole database are consistent 1 m intervals, and 90% are between 0.8 and 1.2 m SRK decided that consequently no compositing was required prior to this initial capping assessment.

Separate histograms and log probability plots were generated for mineralized groups; the groups were defined according to their interpreted mineralogical composition. SRK identified three categories: high-pyrite (≥25% Fe), low-pyrite (<25% Fe), and polymetallic veins (relative low Fe but elevated Ag, Zn and Pb). Review of the histograms and log probability plots identified a small number of significant high-grade outliers, which were subsequently capped. Examples of these outliers are presented in Figure 14‑8 and Figure 14‑9. Consequently, SRK applied the initial top-cuts listed in Table 14.3. SRK notes that the capping thresholds were intentionally conservative, as a second phase of top-cut optimization was envisaged following statistical and visual validation of the initial estimation results, as discussed below.

Following completion of the initial capping assessment, mineralized drill intersections were converted into full intercept composites. For each mineralized intercept, SRK calculated the TT of the interpreted vein, together with the corresponding average capped grade for each analyte (Ag, Au, As, Cu, Fe, Pb and Zn). Metal accumulation values were then calculated as the product of TT and average grade and assigned to a single

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point located at the centre of each interpreted intercept. These metal accumulation and TT values formed the primary estimation variables used throughout the subsequent geostatistical analysis and interpolation.

Following the initial estimation, and during SRK's initial validation phase, additional top cuts based on accumulated grades (grade × true thickness) were applied to individual estimation domains where high-grade samples were observed to unduly influence adjacent blocks, resulting in local smearing of higher grades. These additional top cuts were selected through a review of domain-specific histograms and visual assessment of the estimation results. The selected top cuts are listed in Table 14‑3, an example of the top cut selected for the metal accumulation is presented in Figure 14‑10.

Table 14‑3    Selected top cuts on raw assay data

Zone			Group			Cap Level (g/t or %)															Cumulative Metal Cut (%)
						Au			Ag			Pb			Zn			Cu			Au			Ag			Pb			Zn			Cu
South			EW Polymetallic			NA			1,500			NA			NA			NA			-			-			-			-			-
			Main High Pyrite			NA			4,775			NA			28			NA			-			4			-			1			-
			Main High Pyrite + AgPbZn			3			700			7			15			NA			4			1			1			1			-
			Main Low Pyrite			10			850			NA			15			NA			3			1			-			1			-
			N30 Polymetallic			34			2,565			26			15			NA			4			3			1			0			-
			Secondary High Pyrite			26			545			NA			25			NA			2			2			-			1			-
Central			Main High Pyrite			75			1,085			12			NA			NA			3			1			0			-			-
			Main Low Pyrite			NA			110			10			17			NA			-			2			7			2			-
			NE High Pyrite			NA			515			7			20			NA			-			2			4			4			-
			NS High Pyrite			20			290			8			20			NA			8			1			6			2			-
			NS Low Pyrite			NA			NA			9			7			NA			-			-			1			3			-
North			EW Polymetallic			NA			NA			NA			NA			NA			-			-			-			-			-
			Imarien High Pyrite			NA			NA			NA			NA			NA			-			-			-			-			-
			Main			2			68			NA			NA			NA			2			2			-			-			-
			NE High Pyrite			NA			205			NA			9			NA			-			8			-			2			-
			NS High Pyrite			8			419			NA			9			NA			3			2			-			2			-
			NS Low Pyrite			6			165			NA			9			NA			4			7			-			1			-
			Tizi High Pyrite			20			1,750			NA			NA			NA			3			15			-			-			-

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Figure 14‑8    Histogram of raw Ag g.t sample data for group XX in south

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Figure 14‑9    Histogram of raw Zn % sample data for group XX in central area

Table 14‑4    Selected top cuts on metal accumulation data (grade x thickness) from SRK’s visual validation

Domain			AU			AG			PB			ZN			CU			Fe			TT
1052			0			0			0			0			0.65			NA			3
1091			8			2225			10			15			1.35			NA			5.5
1632			11			265			2.5			NA			0.45			NA			7
2091			70			1000			11.5			35.5			3.15			250			11
2104			19.5			645			6.75			18			1			NA			7.35
2405			1.25			105			5			5			NA			140			4
3101			11.75			265.5			4.15			6.25			0.95			NA			5.5

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3103

10

250

NA

2.5

1.2

NA

6.25

3291

12.5

545

10

16.75

0.87

120

9

3292

6

NA

2.5

NA

0.25

NA

5.65

3301

5.5

1335

5

NA

1.5

NA

6

1004

NA

NA

NA

NA

NA

30

NA

1007

NA

NA

NA

NA

NA

25

NA

1056

NA

NA

NA

NA

NA

100

NA

1090

NA

NA

NA

NA

NA

180

NA

1202

NA

NA

NA

NA

NA

50

NA

1305

NA

NA

NA

NA

NA

100

NA

1311

NA

NA

NA

NA

NA

80

NA

2510

NA

NA

NA

NA

NA

175

NA

3011

NA

NA

NA

NA

NA

40

NA

3121

NA

NA

NA

NA

NA

100

NA

3294

NA

NA

NA

NA

NA

80

NA

3295

NA

NA

NA

NA

NA

200

NA

3301

NA

NA

NA

NA

NA

90

NA

Figure 14‑10    Histogram of raw Ag metal accumulation data for domain 1091 in South

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Figure 14‑11    Histogram of raw Au metal accumulation data for domain 2091 in Central

14.10.3Statistics

A comparison of the uncapped and capped accumulated metal values for Ag, Au, Cu, Fe, Pb and Zn for each estimation domain is presented in Appendix B. SRK notes that the uncapped accumulated metal values incorporate the top-cuts applied to the raw, uncomposited drillhole assay database, as described in 14.10.2.

The distribution of accumulated metal values for most domains approaches a log-normal distribution, examples of which are presented in Figure 14‑12 for Cu and Figure 14‑13 for Ag. Some domains exhibit more complex distributions owing to limited sample support due to the accumulation approach; however, they generally remain consistent with a positively skewed (approximately log-normal) distribution, though in a few cases a normal gaussian distribution is observed.

The coefficient of variation (CoV) for the modelled domains typically ranges between 0.8 and 1.4. The CoV generally decreases where the limited additional top-cuts have been applied to the accumulated metal values. Overall, top-cutting has a minimal effect on the mean accumulated metal grades within most domains. The greatest impact is observed in domains with limited sample support or exceptionally high accumulated Ag values compared to the rest of the domain samples, such as Domains 4001 and 3301. In SRK's opinion, domains with high CoV’s typically require additional drilling to increase the sample support of these domains.

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Figure 14‑12    Example capped metal accumulation histogram for Au in domain 2091

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Figure 14‑13    Example capped metal accumulation histogram for Ag in domain 2091

14.10.4Geostatistical analysis

Experimental variography was undertaken using calculated metal accumulation values (grade × true thickness) for each structural vein group, as defined by the orientation of the interpreted mineralization wireframes in Section 14.9.1.

Log-transformed directional experimental variograms were generated to investigate the principal directions of spatial continuity for Ag, Au, Cu, Pb, Zn and true thickness. The variograms typically have only a few hundred sample pairs, due to the accumulation approach adopted, but has resulted in variograms of moderate to high quality. The principal direction of continuity was clearly identified and corresponds closely with the interpreted plunge of mineralized shoots. Exponential models were fitted to the experimental variograms.

Owing to the limited number of closely spaced downhole sample pairs, the nugget effect was primarily evaluated using omnidirectional variograms together with the shortest lag distances available in the directional variograms.

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Variography was only possible for structural groups N1, N2 Imarirren, N2 Tizi, C1, C2, C4, S1, S2, S3, S4, S5 and S6, owing to the greater quantity of supporting data and the higher confidence in the geological interpretation of these groups. The remaining groups (N3, N4, N5, N6, N7, C3, C5, C6, C7, C8, S7, S8 and S9) contained insufficient data to support meaningful variogram modelling. For these groups, variogram parameters were adopted from the most appropriate analogue group, based on similarities in orientation and geological interpretation. Specifically, the plunge orientation used during estimation was fixed using the corresponding analogue variogram model, whilst the search orientation and variogram geometry were locally adjusted based on the average orientation of the corresponding group of modelled veins using Vulcan LVA function, as discussed in Section 14.10.9.

For the grade continuity analysis, North Group 2 was subdivided into the Tizi and Imarirren vein packages, despite both belonging to the same structural group. This subdivision was undertaken because the two vein packages are spatially distinct and separated by sufficient distance to warrant independent variogram analyses.

The modelled variogram parameters are denoted in Table 14‑6, with example images presented in Figure 14‑14 to Figure 14‑16.

Figure 14‑14    Directional variograms for Pb (grade x thickness) for group C1 (central): (A) downhole variogram; (B) major-axis variogram; (C) semi-major-axis variogram; (D) minor-axis variogram and (E) omnidirectional variogram

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Figure 14‑15    Directional variograms for true thickness (TT) for group N2_Tizzi (north): (A) downhole variogram; (B) major-axis variogram; (C) semi-major-axis variogram; (D) minor-axis variogram and (E) omnidirectional variogram

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Figure 14‑16    Directional variograms for Zn (grade x thickness) for group S2 (south): (A) downhole variogram; (B) major-axis variogram; (C) semi-major-axis variogram; (D) minor-axis variogram and (E) omnidirectional variogram

Table 14‑5    SRK modelled variogram parameters

Group			Element			Nugget			Number of structures			Major angle (Vulcan input) °			Semi Major angle (Vulcan input) °			Minor angle (Vulcan input) °			Sill			Total Sill			Major Range (m)			Semi Major Range (m)			Minor Range (m)
C1			Ag			30%			1			123.9			-22.5			62.8			70%			100%			275			125			50
C1			Au			20%			1			128.5			-13.6			64.2			80%			100%			250			125			50
C1			Cu			25%			1			128.5			-13.6			64.2			75%			100%			110			80			50
C1			Pb			25%			1			128.5			-13.6			64.2			75%			100%			251			150			50
C1			TT			25%			1			128.5			-13.6			64.2			75%			100%			225			130			50
C1			Zn			25%			1			128.5			-13.6			64.2			75%			100%			370			210			50
C2			Ag			25%			1			129.6			-22.5			45.9			75%			100%			200			295			50
C2			Au			25%			1			117.3			-32.8			40.1			75%			100%			200			295			50
C2			Cu			25%			1			129.6			-22.5			45.9			75%			100%			200			265			50
C2			Pb			25%			1			117.3			-32.8			40.1			75%			100%			200			295			50
C2			TT			25%			1			129.6			-22.5			45.9			75%			100%			200			250			50
C2			Zn			25%			1			129.6			-22.5			45.9			75%			100%			200			295			50
C4			Ag			25%			1			141.7			-24.2			51			75%			100%			205			115			50

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Group			Element			Nugget			Number of structures			Major angle (Vulcan input) °			Semi Major angle (Vulcan input) °			Minor angle (Vulcan input) °			Sill			Total Sill			Major Range (m)			Semi Major Range (m)			Minor Range (m)
C4			Au			20%			1			148.2			-16.3			53.3			80%			100%			130			110			50
C4			Cu			20%			1			120.7			-42.1			39.3			80%			100%			150			105			50
C4			Pb			10%			1			148.2			-16.3			53.3			90%			100%			260			135			50
C4			TT			10%			1			125.6			-38.9			42.6			90%			100%			105			135			50
C4			Zn			10%			1			148.2			-16.3			53.3			90%			100%			285			135			50
N1			Ag			25%			1			247.5			21.5			13.1			75%			100%			135			190			20
N1			Au			25%			1			203.6			6.3			24.2			75%			100%			165			85			20
N1			Cu			25%			1			203.6			6.3			24.2			75%			100%			180			70			20
N1			Pb			25%			1			203.6			6.3			24.2			75%			100%			190			105			20
N1			TT			25%			1			203.6			6.3			24.2			75%			100%			205			125			20
N1			Zn			25%			1			208.3			8.3			23.7			75%			100%			175			165			20
N2 Immarien			Ag			25%			1			358.1			24.1			106.5			75%			100%			190			155			20
N2 Immarien			Au			25%			1			356.5			28.9			107.2			75%			100%			195			180			20
N2 Immarien			Cu			25%			1			354.7			33.6			108.1			75%			100%			150			150			20
N2 Immarien			Pb			25%			1			354.7			33.6			108.1			75%			100%			225			220			20
N2 Immarien			TT			25%			1			359.6			19.3			105.9			75%			100%			225			160			20
N2 Immarien			Zn			25%			1			354.7			33.6			108.1			75%			100%			280			140			20
N2_Tizi			Ag			20%			1			1.5			28.9			107.2			80%			100%			180			150			20
N2_Tizi			Au			20%			1			1.5			28.9			107.2			80%			100%			195			180			20
N2_Tizi			Cu			20%			1			314.3			72			147.1			80%			100%			135			125			20
N2_Tizi			Pb			20%			1			1.5			28.9			107.2			80%			100%			330			180			20
N2_Tizi			TT			20%			1			1.5			28.9			107.2			80%			100%			185			150			20
N2_Tizi			Zn			20%			1			1.5			28.9			107.2			80%			100%			345			155			20
S1			Ag			30%			1			329.6			-24.6			-79			70%			100%			165			125			50
S1			Au			20%			1			329.6			-24.6			-79			80%			100%			145			85			50
S1			Cu			20%			1			329.6			-24.6			-79			80%			100%			205			180			50
S1			Pb			20%			1			329.6			-24.6			-79			80%			100%			190			150			50
S1			TT			30%			1			329.6			-24.6			-79			70%			100%			230			95			50
S1			Zn			20%			1			329.6			-24.6			-79			80%			100%			265			175			50
S2			Ag			15%			1			114.3			-69.4			13.5			85%			100%			120			120			50
S2			Au			25%			1			114.3			-69.4			13.5			75%			100%			115			110			50
S2			Cu			20%			1			114.3			-69.4			13.5			80%			100%			115			110			50
S2			Pb			15%			1			114.3			-69.4			13.5			85%			100%			130			110			50
S2			TT			15%			1			114.3			-69.4			13.5			85%			100%			120			95			50
S2			Zn			15%			1			114.3			-69.4			13.5			85%			100%			120			95			50

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Group			Element			Nugget			Number of structures			Major angle (Vulcan input) °			Semi Major angle (Vulcan input) °			Minor angle (Vulcan input) °			Sill			Total Sill			Major Range (m)			Semi Major Range (m)			Minor Range (m)
S3			Ag			25%			1			72.8			-41.6			-18.9			75%			100%			145			105			50
S3			Au			25%			1			79.2			-43.1			-14.5			75%			100%			195			135			50
S3			Cu			30%			1			79.2			-43.1			-14.5			70%			100%			170			120			50
S3			Pb			25%			1			79.2			-43.1			-14.5			75%			100%			150			105			50
S3			TT			25%			1			79.2			-43.1			-14.5			75%			100%			110			105			50
S3			Zn			25%			1			79.2			-43.1			-14.5			75%			100%			110			105			50
S4			Ag			20%			1			354.8			-35.4			-45.3			80%			100%			350			175			50
S4			Au			20%			1			340			-20.3			-52.3			80%			100%			310			175			50
S4			Cu			20%			1			340			-20.3			-52.3			80%			100%			350			175			50
S4			Pb			20%			1			340			-20.3			-52.3			80%			100%			350			175			50
S4			TT			20%			1			340			-20.3			-52.3			80%			100%			165			130			50
S4			Zn			20%			1			340			-20.3			-52.3			80%			100%			305			175			50
S5			Ag			20%			1			316.5			-28.9			- 107.2			80%			100%			190			160			50
S5			Au			20%			1			316.5			-28.9			- 107.2			80%			100%			160			140			50
S5			Cu			20%			1			169.1			-56.8			118.2			80%			100%			130			140			50
S5			Pb			20%			1			316.5			-28.9			- 107.2			80%			100%			190			140			50
S5			TT			20%			1			316.5			-28.9			- 107.2			80%			100%			185			140			50
S5			Zn			20%			1			316.5			-28.9			- 107.2			80%			100%			235			185			50
S6			Ag			20%			1			63.7			-54.7			7.1			80%			100%			200			140			50
S6			Au			20%			1			330.7			-29.5			-78.5			80%			100%			255			240			50
S6			Cu			20%			1			115.2			-35.4			45.3			80%			100%			280			220			50
S6			Pb			20%			1			72.1			-53.8			13.9			80%			100%			350			175			50
S6			TT			20%			1			77.1			-53.8			13.9			80%			100%			320			400			50
S6			Zn			25%			1			72.1			-53.8			13.9			75%			100%			205			145			50

14.10.5Boundary Analysis

Based on previous structural studies, two principal fault sets are recognized to overprint the northwest-southeast-trending mineralized corridor. Northeast-striking (approximately 020°–200°) faults are interpreted to have controlled Au enrichment and a subsequent Zn mineralizing event, whereas younger east-northeast-striking (approximately 070°–250°) faults are associated with Ag-Pb mineralization cross-cut the earlier structural framework.

Based on this, SRK decided to apply soft boundaries to 16 domains in the southern area and 10 domains in the northern area. For the selected domains, a soft-boundary search distance of 30 m × 30 m × 10 m was applied along the major, semi-major and minor axes for all search passes (Table 14‑6), which allows for any close spaced drilling to influence the block estimates in this region. These domains correspond to cross-cutting vein sets that have similar orientations to these two later fault sets. Where drillhole spacing is sufficient, SRK also observed elevated metal grades associated with these intersecting structures, indicating continuity of mineralization across the domain boundaries and supporting the use of soft boundaries for grade estimation.

Hard boundaries were applied between all remaining mineralized domains and between mineralization and waste, where the boundary analysis indicated distinct grade populations. A comparable structural relationship is described in Section 7 and 9.6.

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Table 14‑6    Selected domains with soft boundaries applied

Domain with Soft boundary conditions applied
1091			1090			3291			3010
1501			1201			3001			3011
1502			1202			3002
1503			1203			3003
1504			1040			3004
1505			1003			3005
1041			1004			3009
1502			1007			3191

14.10.6Quantitative Kriging Neighbourhood Analysis (QKNA)

Quantitative Kriging Neighbourhood Analysis (QKNA) was undertaken using Supervisor to support the estimation parameters selected by SRK. The analysis was used to optimize and support the selection of the parent block dimensions, the number of discretization points per block, and the minimum and maximum number of samples used during estimation. Supervisor evaluates these parameters by testing the sensitivity of the kriging efficiency and slope of regression to the parameters.

QKNA was undertaken on the capped Au accumulated metal data (grade × true thickness) for four of the most well informed domains (1091, 2091, 3191 and 3291). The output from Supervisor was used to support the selection of the estimation parameters and parent block dimensions. Based on the QKNA results (Figure 14‑17), a parent block size of 5 m × 15 m × 15 m (X, Y and Z), a minimum of three samples, a maximum of seven samples, a search ellipsoid of 200 m × 150 m × 50 m were selected alongside a discretization of 2 x 2 x 2, though discretization was shown not to be sensitive in this Supervisor analysis.

SRK reviewed the QKNA results in conjunction with the resulting block grade estimates and visual and statistical validation prior to finalising the estimation parameters adopted for the 2026 MRE.

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Figure 14‑17    Example of QKNA results for block sizes (A) and number of samples used in the search (B) for domain 1091 (south)

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14.10.7Block Model and Estimation

A 3D un-rotated block model was generated within Vulcan and sub-blocked by the interpreted mineralization domains, weathering model and the topographic surface.

The block model geometry (extents) and block sizes are presented in Table 14‑7.

Table 14‑7    Boumadine block model framework

Description			Easting (X)			North (Y)			Elevation (Z)
Block Model Origin (Lower left corner)			314500			3473000			0
Mineralization Parent Block Dimension			5			15			15
Waste Parent Block Dimension			90			90			90
Minium Sub-Block Dimension			1			3			3
Rotation			None applied

The parent block dimensions were selected to reflect the average drillhole spacing along strike and down dip, thereby appropriately representing the expected grade variability while maintaining an acceptable degree of smoothing during estimation. The across-strike 5 m block dimension was selected as most modelled domains are less than 5 m wide. However, given that the estimation approach utilized accumulated metal values (grade × true thickness) and a single point for each drillhole, the across-strike block dimension is not considered to be a sensitive parameter. The parent block dimensions within the mineralized domains are approximately one-third of the nominal 50 m infill drillhole spacing, which currently covers just under half of the modelled mineralization wireframes. The block model parameters selected by SRK were supported by the QKNA conducted in Supervisor, as noted above.

SRK recommends that a change-of-support (parent block size) study be undertaken during the next phase of the Project. This should be completed following additional, closer-spaced (infill drilling - which is ongoing currently) to provide more reliable variogram models and to optimize the parent block dimensions and sample selection criteria for future MRE’s.

14.10.8Grade interpolation

Resource estimation was completed within an area encompassing all modelled mineralized domains.

•Metal accumulation (grade × true thickness) and TT were estimated independently within each mineralized domain using OK. Whilst As and Fe was estimated using the same parameters as the other elements using IDW² instead of OK.

•Top-cuts were applied as described in Section 14.10.2.

•Hard boundary conditions were applied during estimation, except for 16 domains in the southern area and 10 domains in the northern area, where soft boundary conditions were employed using a search distance of 30 m × 30 m × 10 m, as described in Section 14.10.5.

•Sub-blocks were assigned the estimated grade of their parent block.

•Sample selection criteria were developed through a combination of visual assessment supported by QKNA undertaken in Supervisor.

•Search ellipsoids were orientated locally to the geometry of each mineralized domain using Vulcan's Local Varying Anisotropy (LVA) functionality (also commonly referred to as dynamic anisotropy). Final block grades were calculated by dividing the accumulation estimate by the TT estimate for each element, in each block.

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14.10.9Estimation Parameters

The selection of the search radii were guided by modelled continuity from the variograms, the typical sample spacing and the QKNA assessment undertaken in Supervisor. The search neighbourhood and sample selection parameters were further refined through a series of initial estimation runs, which were visually reviewed to optimize the estimation parameters. This resulted in the adoption of octant-based sample selection during grade estimation. In some domains, the minimum and maximum sample selection criteria were adjusted to account for limited sample support and to ensure that all domains could be estimated.

The search neighbourhood comprised two successive search passes using the following parameters:

•Pass 1: 200 m x 150 m x 50 m, with a minimum of three and maximum of seven samples, and a maximum of two samples per octant.

•Pass 2: 500 m x 500 m x 500 m, with a minimum of two and maximum of five samples, and a maximum of two samples per octant.

LVA was employed for the estimation of all domains. Three-dimensional surfaces representing the centre of each mineralized vein were constructed and used to calculate the local dip and dip direction throughout the block model. These locally varying orientations were referenced during each estimation pass to align the search ellipsoid with the geometry of each mineralization domain (Figure 14‑18). The same local orientations were also applied to the variogram models; however, the plunge was fixed based on the geostatistical assessment (Table 14‑8).

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Figure 14‑18    Vertical section (looking north) of a modelled domains (red and green) with local search ellipses shown (blue) to demonstrate application of LVA

Table 14‑8    Fixed plunge applied to the estimation group

Group			Element			Azimuth			Dip
C1			All (including TT			128.5			-13.6
C2			All (including TT			117.3			-32.8
C4			All (including TT			148.2			-16.3
N1			All (including TT			203.6			6.3
N2_Immarien			All (including TT			354.7			33.6
N2_Tizi			All (including TT			1.5			28.9
N6_DirectionOnly			All (including TT			118.3			-49
S1			All (including TT			329.6			-24.6
S2			All (including TT			114.3			-69.4
S3			All (including TT			79.2			-43.1
S4			All (including TT			340			-20.3
S5			All (including TT			316.5			-28.9
S6			All (including TT			72.1			-53.8

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14.10.10Density Estimation

A dry in situ density (density) was estimated into the parent block model using IDW². Prior to estimation, SRK reviewed the available density database labelled as SGDRY, to assess its suitability for MRE. During this review, it was observed that the frequency of measured density values within the weakly mineralized and waste sample intervals was substantially lower than the frequency of density measurements collected from the corresponding MS intervals, as discussed in Section 11.6. Consequently, the measured density database alone was considered insufficient to adequately represent the density distribution throughout the deposit.

To reduce this sampling bias, SRK evaluated the relationship between measured dry in situ density (SGDRY) and the calculated total sulphide content (Section 14.4) within the modelled mineralized domains. A broad positive correlation was observed, whereby dry in situ density generally increased with increasing total sulphide content, although the relationship was non-linear.

Several potentially anomalous density measurements were identified during the review. Thirteen samples (shown in orange in Figure 14‑19) were considered to be erroneous, and their SGDRY values were excluded entirely from the density database. The remaining anomalous measurements (shown in red and blue in Figure 14‑19) were retained within the density database but excluded from the regression analysis. Consequently, only the samples shown in green in Figure 14‑19 were used to derive the regression relationship. Figure 14‑20 illustrates the regression fitted to the retained density data, which was subsequently used to estimate density for samples lacking measured SGDRY values. The regression equation is:

Equation 6:

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Figure 14‑19    Coded density database showing the 13 SGDRY samples excluded from the analysis (orange), SGDRY samples retained but excluded from the regression (red and blue) and samples retained for the regressional analysis (green)

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Figure 14‑20    Plot showing regression applied to density database used to determine density when no SGDRY values have been recorded

Based on this relationship, density values were assigned to the coded density database using the following hierarchy:

•Samples identified as "Excluded from regression" (13 samples) were assigned density values calculated using the regression (Figure 14‑20).

•Where measured density values (SGDRY) were available for a given SAMPLEID, the measured values were used directly. Where multiple density measurements existed for the same SAMPLEID, a length-weighted average density was assigned.

•Where no measured density value was available, density was estimated from the calculated total sulphide content using the regression.

Density values were assigned at the assay sample interval (including measured values SGDRY), as the regression relationship was developed using total sulphide values calculated from the assay intervals.

Several drillholes contained neither SGDRY values nor sufficient analytical data to calculate total sulphide content. Consequently, density could not be assigned to these intervals and they remained unpopulated within the coded density database.

A single density composite was subsequently generated for each estimation interval and estimated into the parent block model using IDW². Unlike the grade estimation, density was estimated directly and was not converted to accumulated value (density × true thickness) prior to interpolation.

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In total, 1,848.9 m of sample intervals were assigned measured SGDRY values (based on the Aya sampleID), while 4,554.3 m were assigned density values derived from the regression relationship. Of the intervals for which density could be assigned, approximately 29% were therefore based on measured SGDRY values and 71% on regression-derived values. A further 640.0 m of sample intervals (approximately 9% of the total interval length considered) remained unpopulated because neither measured density nor sufficient analytical data were available to derive a density value. The density search neighbourhood utilized the same set of parameters as the grade estimate (search passes, size of ellipse and number of samples).

A fixed waste density of 2.75 t/m³ was assigned to fresh waste and 2.65 t/m³ to transitional (weathered) waste. These values represent the average measured dry in situ density (SGDRY) within the respective waste domains.

Within the transitional mineralized domains, the estimated density was reduced by 10%. This adjustment was applied because only 17 measured density values were available for the transitional mineralized material, which SRK does not consider sufficient to adequately represent the density characteristics of this domain.

Based on the above observations, SRK recommends that Aya expand its density data collection program, as discussed in Section 14.4. Additional density measurements should be collected across the full weathering profile from a representative range of both mineralized and waste intervals, with the current SRK mineralization wireframes used as a guide for sample selection. This may require additional drilling of the upper portions of the deposit. This additional data should also be used to validate and, where appropriate, refine the sulphide-density regression relationship. In addition, SRK recommends that Aya collect density measurements from drillholes that currently contain no density information in order to improve the spatial coverage of the density database and support future MRE.

The long-term objective should be to estimate density from measured dry in situ density values, thereby removing the reliance on regression-derived estimates.

Following density estimation, block tonnages were calculated by multiplying the estimated dry in situ density by the corresponding parent block volume.

14.10.11Proportion of blocks filled in each search pass

On average, approximately 75% of the mineralized blocks were estimated during the first search pass, with the remaining 25% estimated during the second search pass. The relatively high proportion of blocks estimated during the first search pass is considered appropriate given the current drilling density, the quantity of available sample data across much of the deposit, and the optimized search ellipsoid and sample selection parameters adopted for the estimation.

14.11Block Model Validation

14.11.1Introduction

The block model estimates were validated by completing the following checks:

•global validation by comparison of the composite grade statistics for the back calculated drillhole grades (metal accumulation/thickness) versus block grade assay statistics,

•local validation using visual inspections on sections and plans, viewing the 2D composites metal accumulations and back calculate drillhole grades (metal accumulation /thickness) versus block grades, as well as density values and

•local validation by comparison of average composite grades with average block grades (metal accumulations) and density samples along different directions (swath plots).

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14.11.2Global statistics

The back calculated mean composite grades (metal accumulation/thickness) were compared to the mean, density weighted, estimated block grades to globally validate the estimates. Appendix C compares Ag, Au, Cu, Pb and Fe and Zn values. SRK did not apply declustering to the composite dataset prior to this comparison. In general, SRK considers that a difference of within ±10% between the mean composite grades and the estimated block grades represents an acceptable level of agreement. For some elements within certain domains, differences greater than ±10% were observed. These differences are primarily associated with domains having limited sample support, clustered drilling, rapid variations in the thickness of the modelled mineralization wireframes, or relatively low mean elemental values, where small absolute differences can result in comparatively large percentage differences.

To assess whether these differences represent estimation bias, the block model was further validated through detailed visual comparison of the estimated block grades against the accumulated metal composites and back calculated mean composite grades (metal accumulation/thickness) on plans, sections and in 3D. This review confirmed that the observed differences do not represent a systematic bias in the estimation process.

SRK considers that the global estimates for Ag, Au, Cu, Pb and Zn provide a reasonable representation of the drillhole grades and that no significant global estimation bias has been introduced. As expected, the strongest correlation between composite and block grades is observed within the well-supported estimation domains, whereas domains with limited sample support and/or clustered drilling exhibit a lower degree of correlation.

14.11.3Visual validation

The metal accumulation and back calculated assay grades (meatal accumulation / thickness), coloured by Ag (g/t), Au (g/t), Cu (%), Pb (%) and Zn (%) grade as well as true thickness (m) and density values, were viewed on sections and plans and compared to the block model. These inspections confirmed that the estimates locally conform to the composites. Figure 14‑21 to Figure 14-30 are examples of the visual validation conducted by SRK coloured by an appropriate legend.

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Figure 14‑21    Long section of domain 1031, A = block model vs capped metal accumulation composites (grade x thickness) for Au, B = block model vs back calculated Au g/t grades (metal accumulation /. thickness)

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Figure 14‑22    Long section of domain 1091, A = block model vs capped metal accumulation composites (grade x thickness) for Zn, B = block model vs back calculated Zn % grades (metal accumulation /. thickness)

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Figure 14‑23    Long section of domain 2091, A = block model vs capped metal accumulation composites (grade x thickness) for Ag, B = block model vs back calculated Ag g/t grades (metal accumulation /. thickness)

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Figure 14‑24    Long section of domain 3103, A = block model vs capped metal accumulation composites (grade x thickness) for Pb, B = block model vs back calculated Pb % grades (metal accumulation /. thickness)

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Figure 14‑25    Long section of domain 3193, A = block model vs capped metal accumulation composites (grade x thickness) for Cu, B = block model vs back calculated Cu % grades (metal accumulation /. thickness)

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Figure 14‑26    Long section of domain 3193, A = block model vs true thickness m

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Figure 14‑27    Long section of domain 2091, A = block model vs density composites (tm/3)

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Figure 14‑28    Cross section in the Southern area through 317165 X and 3475490 Y (azimuth 340° dip 90°) showing block model vs back calculated Ag g/t (A), Au g/t (B), Cu % (C), Fe % (D), Pb % (E) and Zn % (F) grades (12 m slicing width)

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Figure 14‑29    Cross section in the Central area through 316911 X and 3477037 Y (azimuth 340° and dip 90°) showing block model vs back calculated Ag g/t (A), Au g/t (B), Cu % (C), Fe % (D), Pb % (E) and Zn % (F) grades (12 m slicing width)

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Figure 14‑30    North south cross section in the Northern area through 315420. X and 3478310 Y (azimuth 0° and dip 90°) showing block model vs back calculated Ag g/t (A), Au g/t (B), Cu % (C), Fe % (D), Pb % (E) and Zn % (F) grades (20 m slicing width)

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14.11.4Swath plots and grade distribution

The block models were also validated by comparing average metal accumulation composite grades and average block grades along different directions (swath plots) and through comparing grade distributions (histograms) of the block model and composite file.

Figure 14‑31 to Figure 14‑34 show examples of swath plots and histograms from some of the best informed domains. The charts show capped metal accumulation composite grades (blue line) against estimated metal accumulation block estimates (grey line), as well nearest neighbour estimate of the metal accumulation (orange line). The numbers of composites used are plotted as blue bars and the model tonnages as red bars on the charts. SRK has not presented the declustered composites in the images below.

Here, and likewise in other domains, the mean composite grades and the mean estimated block grades follow similar trends in all directions. The capped estimates are somewhat smoother than the composite grades, particularly where there are limited samples or very high-grade composites, as demonstrated by the nearest neighbour estimates.

It is SRK’s opinion that the estimate is a reasonable representation of the input composites at the current level of sampling. It should be noted that SRK chose to only cap extremely high grade (raw assays and a metal accumulation values) composites in the domains allowing the representation of the locally higher-grade internal structures.

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Figure 14‑31    Mean metal accumulation (Au) composite Grades (blue line) compared to Block Estimate Grades (orange = OK estimate and grey = nearest neighbour estimate) for domain 1031: swaths along easting (upper left), northing (upper right), elevation (lower left)

Figure 14‑32    Mean metal accumulation (Au) composite Grades (blue line) compared to Block Estimate Grades (orange = OK estimate and grey = nearest neighbour estimate) for domain 1091: swaths along easting (upper left), northing (upper right), elevation (lower left)

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Figure 14‑33    Mean metal accumulation (Ag) composite Grades (blue line) compared to Block Estimate Grades (orange = OK estimate and grey = nearest neighbour estimate) for domain 2091: swaths along easting (upper left), northing (upper right), elevation (lower left)

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Figure 14‑34    Mean metal accumulation (Zn) composite grades (blue line) compared to block estimate grades (orange = OK estimate and grey = nearest neighbour estimate) for domain 3191: swaths along easting (upper left), northing (upper right), elevation (lower left)

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Figure 14‑35    Density composite grades (blue line) compared to block density values (orange = OK estimate) for domain 2091: swaths along easting (upper left), northing (upper right), elevation (lower left)

14.11.5Validation summary

SRK is satisfied that the estimates of density, Au, Ag, Cu, Pb and Zn are a reasonable representation of the input composites at the current level of sampling and that no global biases have been introduced.

14.12Mineral Resource Classification

The 2026 MRE for Boumadine has been classified in accordance with the CIM Definition Standards. Mineral Resource classification is inherently a matter of professional judgement and should consider confidence in the geological interpretation, geological and grade continuity, the quality and quantity of the supporting exploration data, and confidence in the estimated tonnage and grades. These factors should be considered collectively to define contiguous areas of similar confidence.

SRK applied these principles through an integrated assessment of the geological interpretation, drilling density, data quality, estimation quality, and grade continuity. Classification wireframes were developed to define contiguous zones of similar confidence rather than assigning classification on a block-by-block basis. Consequently, some isolated blocks that individually satisfy the classification criteria have been excluded from the final classified volumes to maintain coherent classified volumes.

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The 2026 MRE is supported by drilling completed by Aya and Maya. Historical drilling programs, principally those completed prior to 2018, were excluded from the geological model and 2026 MRE where drillhole locations or sampling quality could not be adequately verified. Review of the QA/QC data indicates that Aya's drilling, sampling and analytical procedures are appropriate for MRE and do not indicate any significant systematic sampling or analytical bias.

Confidence in the geological interpretation is considered moderate to high within the principal mineralized structures, which are supported by modern drilling, surface and underground mapping, historical artisanal workings, historical BRPM underground development, and locally by oriented core measurements, typically were there are four or more individual drill hole intercepts. Geological continuity is locally disrupted by east-west- and N30°-striking faults, resulting in areas of increased uncertainty in the interpreted mineralized volumes. Confidence decreases towards the peripheral vein systems and at depth, where drilling density is limited and structural controls are not well understood.

Confidence in grade continuity is supported by experimental variography and the ongoing infill drilling program. Variogram models of the principal mineralized domains indicate major-axis continuity of approximately 145 m to 250 m, with approximately 70% to 80% of the total grade variance explained within distances of approximately 40 to 60 m. Historical BRPM underground mapping and sampling provide additional qualitative support for the continuity of the principal massive sulphide veins.

The Mineral Resource has been classified principally on the basis of drillhole spacing, geological continuity, grade continuity, and the quality of the estimation support. SRK reviewed the drillhole spacing for each estimation domain in conjunction with the geological interpretation, variography, and the number of informing composites contributing to each estimate. Areas supported by nominal drillhole spacings of approximately 50 m (across a minimum of 3 drillhole fences), where the geological interpretation is well constrained and blocks are informed by multiple drillhole intersections, have been classified as Indicated Mineral Resources. Furthermore, any areas that initially satisfied the criteria for Indicated Mineral Resources within the transitional (weathered) zone were subsequently downgraded to Inferred Mineral Resources. This reflects the limited density data available within the transitional mineralized material, the lack of metallurgical testwork to demonstrate that the material can be processed using the proposed processing route and uncertainty regarding the depth/extent of artisanal surface mining.

Areas supported by nominal drillhole spacings of up to approximately 100 m, where geological and grade continuity remain reasonable but confidence is lower, have been classified as Inferred Mineral Resources, provided that the estimates are informed by multiple drillhole intersections. Any domains supported by three or less drillhole intercepts (typically along a single fence of drilling), or that do not satisfy the Inferred Mineral Resource classification criteria, have not been classified by SRK.

No Measured Mineral Resources have been classified by SRK.

The Inferred Mineral Resource requires additional infill drilling to improve confidence in the geological interpretation, density model, and local grade estimation before consideration can be given to upgrading the classification or supporting detailed mine planning. The resulting Mineral Resource classification is illustrated in Figure 14‑36.

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Figure 14‑36    View of the block model coloured by Mineral Resource classification (orange=Indicated, green=Inferred and grey = unclassified) in relation to the estimation composites (black)

14.13Mining Depletion

The Boumadine deposit has been exploited primarily by underground mining undertaken by BRPM, SODIM and SODECAT between 1986 and early 1992. Also, older exploitation, focussed on near surface mineralization was completed. Records, including maps of the historical underground workings were captured in a local Mine grid. These maps were transformed from the historical local mine grid into the Project UTM coordinate system by GoldMinds Geoservices Inc., providing the basis for incorporating the historical mining information into this MRE (Figure 14‑37).

As discussed in Section 12.2.3, and 12.3, SRK reviewed the georeferenced plans and considers that they provide an appropriate representation of the overall mining geometry for geological interpretation and depletion purposes. However, location discrepancies remain evident between some of the georeferenced underground workings and the interpreted mineralization wireframes. This is considered to reflect the uncertainty associated with the historical mine surveys and the georeferencing process, rather than inaccuracies in the geological model. In several areas, the mined voids do not coincide with the current geological interpretation. It is known that the historical mining targeted the massive sulphide mineralization, whereas the current SRK mineralization wireframes also include adjacent lower-grade mineralization (semi massive and disseminated sulphides). The offsets between the as-built wireframes and the interpreted mineralization could result in portions of mined material remaining within the MRE if depletion were applied solely using the historical mine voids.

SRK also reviewed recent high-resolution satellite imagery to identify areas of historical artisanal mining at surface, which were incorporated into the depletion model where appropriate using a 5m buffer from surface, this buffer also incorporated the QP’s site visit observations (GPS points).

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Figure 14‑37    Underground mining as-builts (stopes = blue and development = black) provided by Aya (location central area)

To ensure that all mineralization believed to have been extracted by underground mining methods was appropriately depleted from the 2026 MRE, SRK constructed additional depletion adjustment wireframes (cookie cutters) in areas where the interpreted mineralized veins and the historical underground workings are (locally) offset (Figure 14‑38). These adjustment solids were designed to fully deplete the mineralized vein targeted by the historical (BRPM, SODIM and SODECAT) underground operations while preserving the overall geological interpretation. The depletion adjustment wireframes were used to trigger local sub-blocking and populate the ugflag field within the block model. This field identifies material interpreted to have been extracted through historical underground mining but which is not necessarily represented within the historical as-built mine void wireframes due to the positional discrepancies discussed above. Aya was informed of the approach and depletion coding prior to finalization of this step.

The following domains have been depleted by SRK:

•1030,1090, 1092, 2004, 2012, 2091, 2104, 2203, 2207, 2314, 3291 and 4002.

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Figure 14‑38    Cross-section looking north at mined as-builts provided by Aya (grey) which is offset from mineralization (mined veins denoted on figure), in relation to SRK depletion adjustment shape (red outline) used to deplete 1030 and 1091

14.14Reconciliation

SRK was provided with limited historical production records, comprising the total tonnes mined only, the grades provided relate to a subset of mining areas specific dates, therefore SRK did not include the grades in the comparison below. No detailed information was provided identifying the specific areas of the deposit mined during individual production periods. Consequently, a direct reconciliation between the historical production records and the depleted MRE for specific areas could not be undertaken. Therefore, SRK compared the total depleted tonnage represented within the block model against the reported historical production (Table 14‑9). In order to check the production tonnage reported by BRPM, Aya provided SRK with the tailings volume, which, using a density of 2.0 t/m3, this equated to 270 Kt, which is only 9 Kt more than the production tonnage reported in Table 14‑9.

As discussed in Section 14.13, there remains a degree of uncertainty regarding the spatial position of the historical underground workings owing to the accuracy of the original mine surveys and the subsequent georeferencing process used to construct the 3D depletion solids. Consequently, it cannot be confirmed that all historical underground mining has been fully represented within the depletion model. In addition, historical artisanal mining at surface has been depleted from the MRE were identified but is not believed to be represented within the reported historical production records, further contributing to differences between the two datasets.

Additional uncertainty arises from the mining methods employed historically. Mining is understood to have been controlled principally using geological mapping, sectional interpretations, underground channel sampling and underground drilling, rather than modern 3D geological and mine planning models. Furthermore, the historical mining is understood to have primarily targeted the massive sulphide mineralization at a relatively high cut-off grade (BRPM reported that the mining cut-off grade was relatively high owing to the marginal economics of the Project at the time), whereas the current SRK mineralization wireframes also include

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adjacent lower-grade mineralization (semi massive and disseminated sulphides) considered to exhibit reasonable geological continuity and economic potential. It is also unknown to what extent mining dilution, overbreak and underbreak affected the historical mining operations.

In total, SRK depleted 485 kt from the block model, representing all material contained within the depletion shapes. The difference between the depleted block model tonnage and the reported historical production is approximately 46% (261 Kt compared with 485 Kt). In SRK's opinion, these differences are considered reasonable given the uncertainties associated with the historical mine surveys, depletion methodology, historical mining practices, changes in the geological interpretation (inclusion of lower grade material within the 2026 wireframes), and the limited historical production information available for review.

SRK cautions the reader against making a direct comparison between the historical production records and the depleted tonnage and grades without considering the uncertainties discussed above. Should underground mining recommence at the Project, SRK recommends that future production stopes are surveyed and routinely reconciled against the geological model and production records. Where safe and practicable, SRK also recommends that Aya resurvey the accessible historical underground workings once these have been dewatered as well as the surface artisanal excavations to further improve the accuracy of the depletion model for future MRE’s and mining studies.

Table 14‑9    Comparison of Mines Production Record and Depleted Block Model

Mined

Mt

Ag g/t%

Au g/t

Cu %

Pb %

Zn%

Production

261,000

N/A

N/A

N/A

N/A

N/A

SRK Model

484,503

82.64

2.89

0.13

0.57

2.46

Difference

223,503

N/A

% Difference

46%

14.15Economic and Technical Input Parameters

Aya developed a Net Smelter Return (NSR) model and cut-off value calculations for both open pit and underground reporting for the Project, which were independently reviewed by SRK. Following several minor amendments, these were adopted for reporting the 2026 Boumadine Mineral Resource Statement. The NSR and associated cut-off values were developed to determine the quantities of material offering "...reasonable prospects for eventual economic extraction", in accordance with CIM requirements, for both open pit and underground mining scenarios.

Aya used reasonable mining and processing assumptions to develop the reporting NSR cut-off values for both open pit and underground Mineral Resource reporting. The NSR reflects the Project location, with a number of input parameters derived from the 2025 Aya PEA, including metallurgical recovery assumptions, mining, processing, treatment charges, penalties, royalites and transportation costs. The NSR cut-off represents the cost of mining, processing and shipping one tonne of mineralized material, including general and administrative (G&A) costs. The parameters used in the NSR calculation are provided in Table 13‑10 to Table 13‑14. The costs in Table 13‑14 have been rounded to a single decimal place.

The metal prices in Table 13‑10 to Table 13‑12 adopted for Mineral Resource reporting are based on the 2025 Aya PEA. In the opinion of the QP, these metal price assumptions are appropriate and consistent with the "reasonable" and "eventual" economic extraction requirements for Mineral Resource reporting.

SRK notes that open pit Mineral Resources take priority over underground Mineral Resources in order to reflect the assumed mining sequence, whereby open pit mining is expected to precede underground mining. Prior to the pit optimization in Whittle software, the block model was regularised to 5 × 5 × 5 m parent blocks. The open pit Mineral Resource Statement was subsequently reported from the in situ sub-block model using a

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marginal NSR cut-off of USD 60/t, based on the technical and economic assumptions listed in Table 13‑14, and includes only classified blocks that occur within, or are intersected by, the conceptual open pit shell that have not been depleted.

In conjunction with the open pit reporting, SRK generated grade shells in Leapfrog to define realistic underground mining targets using an underground NSR cut-off of USD 110/t (Table 13‑14). Areas exceeding this threshold and with a mineralized thickness greater than 1.5 m (assumed minimum mining thickness) were delineated. For mineralized zones less than 1.5 m thick, an NSR × thickness criterion was applied, whereby only blocks with values of 165 (110 (NSR) x 1.5 (thickness)) or greater were considered for reporting. Following this, SRK removed small and isolated grade shells, that were considered unlikely to be mined using underground mining methods, before reporting.

The resulting grade shells, were used to constrain the reporting of the underground Mineral Resource (classified material only). No additional cut-off value was applied directly to the block model. Instead, the entire classified mineralized volume enclosed by the mineable grade shells was reported (waste is excluded), as the shells represent the minimum mining selectivity assumed for the proposed underground mining method at Boumadine. The 2025 Aya PEA states that underground mining will be undertaken using the Avoca mining method, a variation of longhole stoping that uses rockfill as backfill.

SRK recommends that future MRE’s incorporate a Mineable Stope Optimizer (MSO) using parameters representative of the proposed underground mining method.

The calculated NSR $/t is as follows:

The calculation is based on metal prices of 2,800 USD/oz Au, 30 USD/oz Ag, 4.6 USD/lb Cu, 1.0 USD/lb Pb and 1.2 USD/lb Zn.

SRK notes the following opportunities to improve the robustness of the NSR calculation in future versions:

•Royalty has been deducted from the NSR formula using a constant 6.75 USD/t based on the 2025 PEA results. While this is not necessarily incorrect, it is less clear than the usual approach, where royalty would be deducted directly from the concentrate streams, so it remains linked to the value drivers. SRK recommends updating the royalty calculation accordingly in future versions.

•Penalties and minimum deductions are currently included in the treatment charge assumptions. SRK recommends that penalties and deductions are modelled separately from treatment charges, using concentrate grades as inputs, so that the calculation remains linked to changes in concentrate grades.

•Freight costs are currently included in operating costs rather than in the NSR calculation. This treatment is acceptable if applied consistently. However, NSR calculations typically include all realization costs, including freight, because these costs are directly linked to the saleable concentrate mass. SRK recommends moving freight costs into the NSR calculation in future versions.

•Concentrate mass calculations use the original plant feed grades for each concentrate stream. Strictly, Pb concentrate should be based on plant feed grades, Zn concentrate should be based on Pb-circuit tailings grades, and pyrite concentrate should be based on Zn-circuit tailings grades. This would affect the calculated mass pull and therefore shipment costs. The difference is expected to be immaterial for Pb

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and Zn concentrates but may be more relevant for pyrite concentrate. SRK recommends reviewing this basis in future versions, particularly the Fe feed grade used for the pyrite concentrate calculation.

It should be noted that Aya also reports silver (AgEq) and gold equivalent (AuEq) grades. However, these are based solely on metal prices and metallurgical recoveries and do not incorporate the full range of modifying factors included in the NSR calculation (treatment, payability, freight, penalties, royalties and smelter charges). Consequently, the equivalent grades were not used for optimization or Mineral Resource reporting. The calculation of the equivalents follows the same principal as Aya 2025 PEA, and these have been provided for comparative use only. SRK recommends calculating the equivalent grade using NSR factors. This approach will avoid any inconsistency with the NSR framework already in place, making the process more straightforward.

The calculated Ag equivalent (AgEq) is as follows:

The calculated Au equivalent (AuEq) is as follows:

Table 14‑10    NSR Zinc Concentrate Calculation

NSR Zinc Concentrate
Description			Unit			Zn Con Zn
Commodity price			USD/lb			1.2
Concentrator recovery			%			74.7%
Concentrate grade			%			57.4
Payability			%			85.0%
Treatment charge			USD/t dmt			205.5
Freight cost*			USD/t wmt			0

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NSR factor			USD/t per %			14.12
*Freight costs are included in the operating costs and set to zero in the NSR calculation

Table 14‑11    NSR Lead Concentrate Calculation

NSR Lead Concentrate
Description			Unit			Pb Con Pb			Pb Con Cu			Pb Con Au			Pb Con Ag
Commodity price			USD/lb or USD/oz			1			4.6			2,800			30
Concentrator recovery			%			82.0%			73.4%			23.8%			50.4%
Concentrate grade			% Pb			29.6
Payability			%			89.9%			90.0%			95.0%			95.0%
Refining charge			USD/oz									30			1.5
Treatment charge			USD/t dmt			122
Freight cost*			USD/t wmt			0
NSR factor			USD/t per % or g/t			12.87			66.95			20.14			0.44
*Freight costs are included in the operating costs and set to zero in the NSR calculation

Table 14‑12    NSR Pyrite Concentrate Calculation

NSR Pyrite Concentrate
Description			Unit			Py Con Au			Py Con Ag
Commodity price			USD/oz			2,800			30
Fe recovery			%			89.0%
Pyrite concentrate Fe grade			% Fe			38.1
Payability			%			78.0%			78.0%
Freight cost*			USD/t wmt			0
NSR factor			USD/t per g/t			50.14			0.31
*Freight costs are included in the operating costs and set to zero in the NSR calculation

Table 14‑13    Total NSR by Element

Total NSR By Element
Description			Unit			Pb			Zn			Au			Ag			Cu
Zinc concentrate			USD/t per % or g/t						14.12
Lead concentrate			USD/t per % or g/t			12.87						20.14			0.44			66.95
Pyrite concentrate			USD/t per % or g/t									50.14			0.31
Total NSR factor			USD/t per % or g/t			12.87			14.12			70.27			0.75			66.95

Table 14‑14    Cut-off value calculation

Parameters			Unit			USD/t
Processing rate per annum			Mtpa			2.9
Cost assumptions
Government rent (mining tax)			USD/t			0.3
General and administrative			USD/t			6
Mining - Open pit
Open pit grade control, drilling and assay			USD/t			1.1
Mineralised material mining cost			USD/t			2
Waste mining cost			USD/t			1.5
Surface haul and rehandle			USD/t			2

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Pit Angle			°			47
Mining - Underground
Underground stoping, including backfill			USD/t			35
Underground operating development			USD/t			13.3
Underground owners costs			USD/t			5.3
Surface haul and rehandle			USD/t			2
Processing			USD/t			17.3
Concentrates Shipment
Lead concentrate			USD/t			1.5
Zinc concentrate			USD/t			1.9
Pyrite concentrate			USD/t			28.2
Water + TSF			USD/t			0.4
Cut-off Values
Open pit costs			USD/t			59.1
Open pit cut-off value for MRE reporting			USD/t			60
Underground costs			USD/t			111.2
Underground cut-off value for MRE reporting			USD/t			110

* costs in the above have been rounded to a single decimal place

14.16Mineral Resource Statement

The SRK 2026 Mineral Resource Statement (MRS) for the Boumadine deposit is presented in Table 14‑15. The 2026 MRE is reported and classified in accordance with the CIM Definition Standards for Mineral Resources and Mineral Reserves (May 2014) and NI43-101 Standards of Disclosure for Mineral Projects (May 2016).

The 2026 MRE has an effective date of 28 February 2026, and has been depleted to reflect the current understanding of the historic mining. The Mineral Resources are limited to classified material contained within conceptual open pit shell reported above the NSR reporting cut-off of 60 USD/t and, below the pit shell, to classified material contained within grade shells (following the removal of isolated shapes) based on a 110 USD/t NSR cut-off, that satisfy a 1.5 m minimum minable thickness. The Mineral Resources are reported within the Boumadine Mining Licence on an in-situ, undiluted basis and amount to:

•Indicated Mineral Resources (open pit and underground) of some 8,582 kt, at a mean grade of 121.7 g/t Ag, 1.9 g/t Au, 0.09 % Cu, 1.01 % Pb and 2.19 % Zn; and

•Inferred Mineral Resources (open pit and underground) of some 45,365 kt, at a mean grade of 56.8 g/t Ag, 1.8 g/t Au, 0.08 % Cu, 0.58 % Pb and 1.33 % Zn.

The reader is cautioned that the results from this study are used solely for the purpose of testing the “reasonable prospects for economic extraction” and do not represent an attempt to estimate Mineral Reserves. The results are used as a guide to assist in the preparation of a MRS.

The Company’s Boumadine Mining Licence permit expired in May 2026. Aya have informed SRK that they are currently in the processes of renewing the Licence (no change to the extent or conditions) for an additional 10 years. Whilst the licence renewal approval is currently pending, the Company has informed SRK that they see no reason why this will not be successful.

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Table 14‑15    SRK Mineral Resource Statement for the Boumadine Project, Morrocco, as at an effective date of 28 February 2026

			Cut-off (NSR USD/t)			Tonnes (Kt)			Average Grade																					Contained Metal
									Ag (g/t)			Au (g/t)			Cu (%)			Pb (%)			Zn (%)			AgEq (g/t)			AuEq (g/t)			Ag (koz)			Au (koz)			Cu (kt)			Pb (kt)			Zn (kt)			AgEq (koz)			AuEq (koz)
Pit-Constrained
Indicated			60			6,639			115.5			2.2			0.1			0.96			2.21			390.8			4.2			24,653			463			6			64			147			83,423			896
Inferred			60			19,497			57.8			2			0.07			0.62			1.54			298.7			3.2			36,218			1,280			14			120			301			187,258			2,011
Underground
Indicated			110			1,943			143			1.1			0.06			1.19			2.11			319.6			3.4			8,931			69			1			23			41			19,963			214
Inferred			110			25,868			56			1.6			0.08			0.56			1.17			251.9			2.7			46,605			1,372			21			145			303			209,475			2,249
Total
Indicated			60/110			8,582			121.7			1.9			0.09			1.01			2.19			374.7			4			33,585			532			8			87			188			103,386			1,110
Inferred			60/110			45,365			56.8			1.8			0.08			0.58			1.33			272.0			2.9			82,823			2,653			35			265			604			396,733			4,260

In reporting the Mineral Resource Statements, SRK notes the following:

•The Mineral Resources have an effective date of 28 February 2026. Mr Guy Dishaw, P.Geo., is the Qualified Person (QP) responsible for the Mineral Resource Estimate in accordance with the Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Definition Standards for Mineral Resources and Mineral Reserves (2014) and CIM Estimation of Mineral Resources and Mineral Reserves Best Practice Guidelines (2019). Mr Dishaw is a member of the Association of Professional Engineers and Geoscientists of Saskatchewan.

•The Mineral Resource Estimate was prepared by a team of consultants from SRK under the supervision of the QP.

•Mineral Resources are reported in situ and undiluted and are constrained within the Boumadine Mining Licence.

•Mineral Resources have been depleted to reflect SRK's current understanding of historical underground mining completed up to mine closure in 1992 as well as surficial artisanal workings.

•Commodity prices of US$2,800/oz Au, US$30/oz Ag, US$4.60/lb Cu, US$1.00/lb Pb and US$1.20/lb Zn were used to establish reasonable prospects for eventual economic extraction.

•The NSR was calculated as: NSR (US$/t) = (Pb (%) × 12.87) + (Zn (%) × 14.12) + (Au (g/t) × 70.27) + (Ag (g/t) × 0.75) + (Cu (%) × 66.95) − 6.75.

•Silver equivalent (AgEq) and gold equivalent (AuEq) grades are based solely on metal prices and metallurgical recoveries and do not incorporate the full range of factors included in the NSR calculation, including treatment charges, payabilities, penalties, royalties and other smelter charges. Consequently, equivalent grades were not used for optimization or application of reporting cut-offs and are provided for reference only.

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•AgEq = Ag (g/t) + (Au (g/t) × Au price/gram × Au recovery) / (Ag price/gram × Ag recovery) + Zn (%) × Zn price/lb × Zn recovery / (Ag price/gram × Ag recovery) × 685.7147973 + Pb (%) × Pb price/lb × Pb recovery / (Ag price/gram × Ag recovery) × 685.7147973 + Cu (%) × Cu price/lb × Cu recovery / (Ag price/gram × Ag recovery) × 685.7147973.

•AuEq = Au (g/t) + (Ag (g/t) × Ag price/gram × Ag recovery) / (Au price/gram × Au recovery) + Zn (%) × Zn price/lb × Zn recovery / (Au price/gram × Au recovery) × 685.7147973 + Pb (%) × Pb price/lb × Pb recovery / (Au price/gram × Au recovery) × 685.7147973 + Cu (%) × Cu price/lb × Cu recovery / (Au price/gram × Au recovery) × 685.7147973.

•Open-pit Mineral Resources are reported within an optimized pit shell using an NSR cut-off value of US$60/t. The optimization and reporting assumptions include mineralised material mining costs of US$2.0/t, waste mining costs of US$1.5/t. Drilling and blasting costs (grade control) US$1.1/t, US$2.0/t surface haulage and rehandling costs, processing and shipping costs of US$49/t, G&A costs of US$6/t and an overall pit slope angle of 47°.

•Underground Mineral Resources are reported using a rounded NSR cut-off value of US$110/t, based on total underground mining costs of US$55.6/t, processing and shipping costs of US$48.9/t and G&A costs of US$6.7/t. The underground Mineral Resources demonstrate sufficient continuity suitable for long-hole underground mining methods. Reporting shapes apply a minimum true mining width of 1.5 m, with isolated or discontinuous volumes considered unlikely to support potential underground extraction excluded from the reported Mineral Resource.

•Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability.

•The QP is unaware of any known environmental, permitting, legal, title, taxation, socio-economic, marketing, political or other relevant factors that could materially affect the Mineral Resource Estimate.

•Tonnages are reported in metric units and grades are reported in grams per tonne (g/t) for Au and Ag and percent (%) for Cu, Pb and Zn. Tonnages, grades and contained metal quantities have been rounded appropriately. Rounding may result in apparent differences between totals and the sums of individual values; such differences are not considered material.

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14.17Sensitivity Analysis

The Mineral Resources of the Boumadine Project are sensitive to the selection of the reporting cut-off value. To illustrate this sensitivity, Figure 14‑39 to Figure 14‑40 present grade-tonnage curves for the combined Indicated and Inferred Mineral Resources for the open pit and underground mining scenarios. The reader is cautioned that the figures presented in these charts do not constitute a MRS and are provided solely to demonstrate the sensitivity of the block model estimates to the selection of the NSR cut-off value. Furthermore, the underground grade-tonnage curve includes all material reported beneath the optimized open pit shell and does not take into account mineable shapes or minimum mining thickness.

Figure 14‑39    Boumadine Grade Tonnage Curves for Open Pit Indicated and Inferred Material

Figure 13‑40    Boumadine Grade Tonnage Curves for Underground Indicated and Inferred Material

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14.18Aya Peer Review

Throughout the project, Aya undertook a peer review of SRK's 2026 MRE at the key stages (upon completion of the wireframing, grade and density estimation as well as SRK’s approach to depletion). All key decisions were discussed during the weekly project calls including SRK’s classification and Mineral Resource reporting approach. Where appropriate, SRK incorporated Aya's feedback, with adjustments being made based on their Project knowledge and constructive feedback.

14.19Comparison to Previous Estimates

As stated in Section 6.8, four previous MRE’s have been completed for the Boumadine Project, with the most recent reported by Aya in March 2025 (2025 MRE), which was in accordance with NI 43-101 (Table 6‑4).

The methodologies used to define the 2026 MRE differ from those used for the 2025 MRE. The principal differences include:

•Additional data and geological modelling: SRK remodelled the Boumadine deposit using all available drilling data (historical BRPM and 2017 Maya drilling was excluded). The 2026 geological interpretation considered an additional 453 drillholes compared to the 2025 MRE, associated with Mineral Resource definition and extensions to the known mineralization. The significant increase in available drilling data is considered to be one of the principal reasons for the differences between the two estimates. The additional drilling comprised both infill and exploration drilling, resulting in upgrades to portions of the previously classified Mineral Resource (to inferred and or indicated) and extensions to the known mineralization. Both the Aya 2025 and SRK 2026 mineralization wireframes were modelled using the Leapfrog vein modelling tool, and both estimates excluded the same drillholes from the geological modelling and grade estimate. The 2026 interpretation is built upon the previous geological models and incorporated the additional drilling, previous wireframe interpretations, depletion shapes and underground geological mapping. SRK adopted a more conservative approach to the geological modelling of individual domains, locally excluding wider intervals of waste and/or low-grade material that had been included within the 2025 wireframes. In some cases, individual domains interpreted in the 2025 MRE were subdivided by SRK into multiple domains based on the updated geological, structural interpretation and their respective orientations. Both the 2025 and 2026 wireframes considered the structural interpretations developed by Gaboury, satellite imagery and structural measurements obtained from oriented drill core. As part of the 2026 study, SRK modelled 158 mineralized domains across the Project, comprising 57 domains in the southern area, 58 in the central area and 43 in the northern area. In comparison, the 2025 MRE comprised 45 modelled domains in total, including 13 in the southern area, 13 in the central area and 19 in the northern area. Some of the additional domains modelled by SRK reflect new geological information provided by the additional infill and exploration drilling that was either unavailable or poorly supported at the time of the 2025 MRE.

•Grade estimation: The 2026 grade estimation methodology differs from that used for the 2025 MRE, principally through the use of a 2D metal accumulation approach rather than 3D grade estimation adopted by Aya in 2025. SRK estimated metal accumulation (grade × true thickness) using ordinary kriging (OK) and octant-based sample selection for Ag, Au, Cu, Pb and Zn. In comparison, the 2025 MRE estimated grades directly from 1 m composites using inverse distance weighting and did not use ordinary kriging or octant-based sample selection.

•Treatment of high-grade outliers: Both the SRK 2026 and Aya 2025 estimation approaches applied high-grade caps to limit the influence of anomalously high-grade samples, albeit using different thresholds and methodologies. For the 2026 MRE, SRK initially applied capping to the raw assay grades prior to calculating metal accumulation. Capping thresholds were assessed using histograms and log-probability

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plots for three mineralization categories: high-pyrite (≥25% Fe), low-pyrite (<25% Fe), and polymetallic veins (relatively low Fe but elevated Ag, Zn and Pb). SRK subsequently applied additional capping to metal accumulation values within selected domains where further control of high-grade samples was considered necessary following visual and statistical validation of the model. In comparison, the 2025 MRE applied capping to the 1 m composites, with thresholds selected by Aya using histograms and log-probability plots for specific domains (if warranted).

•Density estimation: The density estimation methodology was revised for the 2026 MRE. SRK estimated density using IDW² from a single composite (across strike intersection). The estimation dataset comprised measured dry in situ density values supplemented, where required, by proxy density values derived from a regression relationship between measured dry in situ density and calculated total sulphide content. In comparison, the 2025 MRE applied fixed density values based on the average measured densities of the modelled mineralized wireframes.

•Depletion: The 2026 MRE was depleted for historical underground mining and surface artisanal workings, whereas the 2025 MRE was not depleted. As discussed above in Section 14.13, where mining is known to have occurred but an offset exists between the modelled mineralization and the available underground as-built wireframes, SRK created additional depletion shapes to account for the interpreted mined material. SRK also depleted surface artisanal workings identified during the QP site visit and from satellite imagery. These depletion shapes were used to ensure that known or interpreted historically mined material was excluded from the 2026 MRE to the best knowledge of the QP.

•Adjustments to the technical and economic parameters used for Mineral Resource reporting: The assumptions and calculations used to derive the 2026 open pit and underground NSR cut-off values were reviewed by SRK as part of the current study. In several cases, Aya made amendments following SRK's review. These included, but were not limited to, the use of concentrate recoveries and grades as inputs to the NSR calculation, revision of the Pb treatment charge to remove previous double counting, and adjustments to concentrate mass pull and associated shipment costs. The 2026 NSR calculation also incorporates updated metal price assumptions. In addition, the processing cost assumptions used to support both open pit and underground Mineral Resource reporting differed from those applied in 2025. The 2026 open pit optimization also incorporated revised mineralized material and waste mining costs and a slightly different overall pit slope assumption. Collectively, these changes resulted in differences in the technical and economic parameters used to demonstrate reasonable prospects for eventual economic extraction and to report the 2026 MRE.

•Mineral Resource reporting: For the 2025 MRE, open pit Mineral Resources were reported within an optimized pit shell above an NSR cut-off value of USD 95/t, while underground Mineral Resources were reported beneath the pit shell above an NSR cut-off value of USD 125/t. Aya did not undertake a stope optimization exercise or exclude volumes that did not form potentially mineable underground shapes or consider any minimum minable thicknesses in reporting the 2025 underground Mineral Resources. The 2025 reporting parameters were derived from the pit optimization and NSR calculations completed at the time. For the 2026 MRE, SRK reported open pit Mineral Resources within an optimized pit shell using an NSR cut-off value of USD 60/t, while underground Mineral Resources were assessed using an NSR cut-off value of USD 110/t. For underground reporting, SRK generated grade shells in Leapfrog, based on block centroid values, to define realistic potential underground mining targets. These grade shells were used to identify areas where the mineralized thickness exceeded 1.5 m, representing the assumed minimum mining thickness. For mineralized zones less than 1.5 m thick where elevated grades were present, an NSR × thickness criterion was applied. SRK subsequently removed small and isolated volumes considered unlikely to be mined using the proposed underground mining methods. This approach provides a more explicit assessment of the potentially mineable geometry of the underground Mineral Resources than that applied in the 2025 MRE.

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Overall, the 2026 Mineral Resource Statement (Table 14‑15) reports higher tonnages than the 2025 MRE. Indicated Mineral Resources increased by approximately 66%, from 5,169 kt to 8,582 kt, while Inferred Mineral Resources increased by approximately 55%, from 29,196 kt to 45,365 kt. The average Ag grade of the Indicated Mineral Resources increased by approximately 34%, from 90.8 g/t to 121.7 g/t, while the Inferred Ag grade decreased by approximately 31%, from 81.8 g/t to 56.8 g/t. Average Au grades decreased by approximately 31% for Indicated Mineral Resources (2.8 g/t to 1.9 g/t) and by approximately 31% for Inferred Mineral Resources (2.6 g/t to 1.8 g/t). Average Zn grades decreased by approximately 22% for Indicated Mineral Resources (2.8% to 2.19%) and by approximately 37% for Inferred Mineral Resources (2.11% to 1.33%). The average Cu grade decreased by approximately 26% for Indicated Mineral Resources (0.12% to 0.09%), while the Inferred grade remained unchanged. The average Pb grade increased by approximately 19% for Indicated Mineral Resources (0.85% to 1.01%) but decreased by approximately 29% for Inferred Mineral Resources (0.82% to 0.58%).

Given the differences in the quantity of available drilling data (an additional 453 drillholes), geological interpretation, estimation methodology, density estimation, depletion, the approach used to demonstrate reasonable prospects for eventual economic extraction, and the technical and economic input parameters, SRK considers that the changes between the 2025 and 2026 Mineral Resource estimates cannot be attributed to any single factor, but a combination of the above.

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15.0MINERAL RESERVE ESTIMATES

This section is not applicable to this Report.

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16.0MINING METHODS

16.1General Description of the Mining Project

The Boumadine deposit is a large polymetallic mineralized system containing Au, Ag, Zn, and Pb, extending over 5 km, with historic mining infrastructure in the central and southern zones. Mineralization at Boumadine begins at the surface in some occurrences and generally takes the form of vertical or subvertical structures. A combined open-pit (“OP”) and underground (“UG”) mining operations are planned for Boumadine mineralized zones.

The mine plan presented in this section is based on the Mineral Resource estimate described in Section 14 and includes Inferred Mineral Resources. Inferred Mineral Resources are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorised as Mineral Reserves. There is no certainty that the results of this Preliminary Economic Assessment will be realised. No Mineral Reserves have been estimated for the Project.

16.2Geotechnical

16.2.1Open Pit

16.2.1.1Open Pit Slope Design

Geotechnical evaluations of the Boumadine pits have been completed by empirical evaluation to produce slope design constraints domained by slope dip direction (see RockEng report #24017-102 for complete documentation of geotechnical site characterization and analyses). The pit slope design guidance may be applied to varying pit shell geometries with reliance on slope orientation (rather than geographical location).

Overall slope angle (OSA) constraints are summarized in Table 16-1. It is noted that these are maximum allowable OSA values; slopes may be limited to a flatter angle once benches, ramps, slope breaks etc. are accounted for in geometry.

Table 16-1    Summary of OSA recommendations.

Slope Dip Direction			Overall Slope Angle (OSA)
			Slope Height up to 380m			Slope Height up to 220m			Slope Height up to 100m
5			48			52			60
65			38			42			48
125			48			52			60
185			48			52			60
245			48			52			60
305			48			52			59

Further to the OSA summarized above, pit slope designs are constrained by the following:

Adverse jointing is expected at Boumadine in east-northeast dipping slopes due to the presence of a prominent ENE dipping joint set. Here, the inter ramp slope angle (IRA) should not exceed the dip of the adverse joint (48˚).

Minimum catch bench width are defined as follows (Ryan & Pryor, 2000 in Read & Stacey, 2009):

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Catch Bench Width (m) = 0.2 * Bench Height + 4.5m

•Where overburden is present, it will be stripped and laid back at a 2H:1V slope. A bench between the toe of the overburden and the crest of rock should be left around the pit perimeter to prevent overburden from running into the pit (bench width should be equal to overburden depth).

•The first 2 benches will be single bench height to account for the likely occurrence of poorer ground conditions near surface (weathering horizon).

•Below the initial two benches in rock, 20 m high double benches are geotechnically feasible.

•Geotechnical slope breaks will be established every 100 vertical metres; preliminary design will consider geotechnical slope breaks to be not less than double the width of a typical catch bench.

16.2.2Underground

Geotechnical evaluations of the Boumadine underground mine have been completed by empirical evaluations (see RockEng report #24017-102 for complete documentation of geotechnical site characterization and analyses).

16.2.2.1Mining Method Selection

Longhole open stoping is a geotechnically sound mining method for the Boumadine underground. Based on deposit body geometry, longitudinal open stoping and/or Avoca style stoping is deemed amenable from a geotechnical perspective. Longitudinal stoping provides operational flexibility to adjust stope panel lengths, thus there is an opportunity for optimizing design for local variances (good or bad) in ground conditions. This is particularly important where the deposit body is narrow as production performance in narrow zones is likely to be more sensitive to HW and FW dilution.

At relatively shallow depths, low stress conditions allow for operational flexibility with mine sequencing. At this phase of study, one key aspect of sequencing that has been identified pertains to inter-lens pillar stability and adjacent stope performance. The lead/lag between adjacent lenses is minimized as an opportunity to mitigate loss of confinement.

16.2.2.2Open Stope Design and Overbreak Potential

Stope stability has been evaluated by empirical methods considering vertical stope sidewalls, HW and FW faces dipping ENE at ~85° (based on modelled stope geometries), and horizontal stope backs. It is assumed that both sulphides and Tuff may be exposed in stope faces. The recommended hydraulic radii limitations are provided in Table 2. It has been concluded that 20 m level spacing (25 m stope height), with 20 m strike lengths are well suited for Boumadine underground. Larger stopes may be achieved if cable bolting is implemented. At these dimensions, with typical HW to FW spans, stope backs and endwalls are expected to perform well, with stope back cable bolting required rarely (where HR exceeds 4.2). The predicted overbreak (for 20 m strike x 20 m level spacing) is roughly 0.5 m for stope HWs and blast damage only for stope FWs.

Table 16-2    Recommended stope sizing constraints by stope face and rock type.

Stope Face			Unsupported HR			Supported HR
HW/FW			5.6			7.5
Back			4.2			6.2
Endwalls			7.5			-

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16.2.2.3Potential Overbreak

HW and FW overbreak is an important metric for dilution planning. An estimate of stope HW and FW overbreak has been assessed using the empirical method by Clark (1998). For hydraulic radii of 5.6 (20 m strike length, with 20 m level spacing) the predicted overbreak is roughly 0.5 m for stope HWs and blast damage only for stope FWs.

16.2.2.4Pillar Stability

Inter-lens Stope Pillars

Inter-lens pillars exist between stacked (HW to FW) mineralized zones. The Boumadine underground mine plans intend to utilize cemented rock fill, and so the largest mining risk associated with inter-lens stope pillar stability is dilution and oversize entering open stopes (as opposed to global mine stability). The following pillar dimensioning constraints have been established by empirical methods.

•At W:H = 0.3 (6 m for a 20 m stope strike length), total pillar loss predicted for HW to FW sequencing is 2.3 m, and for FW to HW sequencing is 2.2 m.

•At W:H = 0.4 (8 m pillar for a 20 m stope strike length), total pillar loss for HW to FW sequencing is 1.8 m, and for FW to HW sequencing is 1.5 m.

•At W:H = 0.5 (10 m pillar for a 20 m stope strike length), total pillar loss for HW to FW sequencing is 1.1 m, and for FW to HW sequencing is 1.0 m.

Table 16-3 Pillar Dilution Based by Mining Sequence.

Pillar Dimensions						0.5W:H			0.4W:H			0.3W:H
HW to FW advance			ELOS 1st Stope (m)			0.3			0.3			0.3
			ELOS 2nd Stope (m)			0.8			1.5			˃ 2.
			Total Pillar Dilution (m)			1.1			1.8			˃2.3
FW to HW advance			ELOS 1st Stope (m)			0.7			0.7			0.7
			ELOS 2nd Stope (m)			0.3			0.8			1.5
			Total Pillar Dilution (m)			1.0			1.5			2.2

Source: RockEng

Rib & Sill Pillars

Rib pillar dimensioning will rely on a 1:0.7 width (rib pillar horizontal thickness) to height (HW to FW span) aspect ratio with a minimum dimension of 3 m. Permanent sill pillar dimensioning relies on a 1:1 width (vertical dimension) to height (HW to FW span) aspect ratio with a minimum dimension of 5 m.

Crown Pillar

Based on empirical crown pillar evaluations, a 20 m pillar (in rock, ignoring overburden thickness) is deemed appropriate for stable crown pillar conditions during active mining operations. This dimension relies on the assumption that backfill will be utilized to manage long term stability for mine closure. Further, it is not unreasonable to expect that, with appropriate engineering and planning, full crown pillar recovery may be achieved (assuming surface disturbance is permitted, and environmental risks are tolerably low).

16.2.2.5Infrastructure Siting

Given the relatively shallow mining depth, the geotechnical risks associated with stress induced damage to permanent infrastructure are considered low; thus, allowing for reasonable design flexibility, general guidance

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is as follows: cross-cut lengths ensure that haulage drive stand-offs are at least 15 m from the deposit body contact; either HW or FW infrastructure placement is acceptable given the near-vertical nature of the deposit; lateral development lay-outs will ensure that wide spans and intersections are not stacked level-to-level.

16.2.2.6Ground Support

Standard lateral development (4 to 6 m spans) will be supported by 2.4 m rebar at a 1.25 x 1.25 m square pattern applied to the back and walls, with welded wire mesh in the back and shoulders. Cable bolts will be utilized in intersections and wide spans wit bolt lengths equal to ½ the span, at a 2 m square pattern. It is assumed that 2% of development will require shotcrete.

16.3Hydrogeology

No site-specific hydrogeological investigation has been completed to date. For the purposes of this PEA, groundwater inflows have not been estimated: the mine dewatering system described in Section 16.6 has been sized on a level-by-level basis from typical operating practice. CCE Mining recommends that a hydrogeological programme comprising packer testing, pumping tests and the installation of monitoring wells be completed, and that a numerical groundwater model be developed, to support the next stage of study.

16.4Boumadine Mining Method

16.4.1Mining Method Selection

This section details the assumptions and procedures used to develop a mining plan for the project. The Boumadine deposit extends from the surface to depths well beyond those that can be mined using an open-pit (OP) method alone. It was decided from the outset that the mining method would be a combination of open-pit (OP) and underground (UG) mining. Initially, open-pit design shells were created using Whittle software, and underground stope designs were created using Deswik Stope Optimizer (DSO). This initial design allowed for a preliminary assessment of the interaction between open-pit and underground mining, and for the development of an open-pit mining plan that would recover as much high-grade mineralized material as possible at a reasonable cost before the underground method became more profitable.

Historical exploration activities, including underground development work, were conducted at the deposit in the 1980s and 1990s; therefore, it will be necessary in the future to locate and map these tunnels so that they can be considered in the detailed design and planning of the mine.

16.4.2NSR Cut-offs determination

To assess the Boumadine project, Aya used a Net Smelter Return (NSR) cut-off to assess the economic viability of mining one metric ton of diluted mineralized material. Viability is determined by comparing the material’s NSR value to the applicable NSR cut-off values. An NSR threshold represents the on-site cost associated with extracting and processing one metric ton of mineralized material, including general and administrative (G&A) expenses.

The NSR cut-offs incorporate mining, processing, and G&A costs based on a preliminary cost model, which has since been adjusted by the operating cost estimates developed in Section 21.

Table 16-4 presents the calculations used to determine the NSR Cut-off thresholds applied in this preliminary economic assessment (PEA) study for open-pit mining, underground mining, and underground development.

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Table 16‑3    NSR Cut-off Calculations General Assumptions

Area			Unit			Value			Source
Metal Price
Lead (Pb)			US$/lb			1			Aya
Copper (Cu)			US$/lb			4			Aya
Zinc (Zn)			US$/lb			1.2			Aya
Gold (Au)			US$/oz			$	2,800.00		Aya
Silver (Ag)			US$/oz			30			Aya
Financial Assumptions
Royalties			%			3
Open-pit Mining
Mining			USD/t			$4.60			Aya
Processing Cost			USD/t			$50.35			Aya
G&A			USD/t			$6.30			Aya
Total (OP NSR Cutoff)			USD/t			$61.25
Waste Mining			USD/t			$2.00			Aya
Underground Mining
Mining			USD/t			$68.13			Aya
Processing Cost			USD/t			$50.35			Aya
G&A			USD/t			$6.30			Aya
Total			USD/t			$124.78
Underground Development
Mining			USD/t			$20.63			Aya
Processing Cost			USD/t			$50.35			Aya
G&A			USD/t			-
Total			USD/t			$70.98
NB - assumes development is required by block - decision to sent to mill vs waste

16.5Open Pit Mining

16.5.1Whittle Open Pit optimization

GEOVIA Whittle (2026) software was used to carry out the pit optimization process for the Boumadine pit. Using the Lerchs-Grossmann (LG) algorithm integrated into Whittle, nested shells were generated, and the optimal pit shell was selected for each area of the deposit. These pit shells were then used for mine design and, ultimately, for the scheduling process.

Estimated costs and cut-off grades were provided by AYA and verified by the QP. These costs were compared to benchmark costs from similar projects before being incorporated into Whittle.

Table 16-5 presents the economic and mining data used for Whittle pit optimization. Revenue for each block was provided by the calculated NSR value.

Table 16‑4    Whittle Parameters

Parameters			Unit			Value
General Assumption
Processing Plant Throughput			tonne/day			8 000
Processing Plant Throughput			tonne/year			2 880 000

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Working Day			day/year			360
Mining Recovery			%			100%
Mining Dilution			%			0%
Mining Costs
Whittle Mining Reference Costs			US$/t			2.00
Processing Costs
Crushing/Grinding/Flotation			US$/t of feed			20.00
Concentrates shipping			US$/t of feed			30.00
Industrial Water/TSF			US$/t of feed			0.35
Additional Mineralized Mining Costs			US$/t			2.60
G & A			US$/t			6.30
Whittle processing Costs			US$/t			59.25

16.5.2Whittle Pit Shell Generation

Pit Shell Generation was first performed on the overall deposit to determine the possible pits to be evaluated. This exercise resulted in the identification of 4 distinct areas with potential pits: 2 in the south (one of which was not economical using open-pit mining method), one in the center, and one in the north. The initial block model was then modified to create four separate block models, each constrained to the boundaries of the largest pits generated during the calculation. Each pit was then optimized individually, allowing for a more precise analysis tailored to the geological and economic characteristics of each area.

The following paragraphs will describe the methodology and optimization parameters used for the central pit. A similar method and parameters were used for the other pits.

The first step was to re-block the 5m x 5m x 5m block models to 15m x 15m x 10 m (X, Y, Z). Figure 16-1 shows the Whittle block size adjustment.

Figure 16-1    Whittle Re-Blocked Model Parameters

Next, the overall slope angles were determined in accordance with geotechnical recommendations, with variations based on the depth of the pits. Figure 16-2 presents the parameters used in Whittle in the first iteration using conservative OSA numbers. After the first iteration, pit depth was determined and OSA was adjusted to the most aggressive recommendation allowed, then Whittle was re-run.

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Figure 16-2    Whittle Initial Slope Set Settings

For each pit, optimized pit shells were generated using revenue factors (RF) ranging from 0.25 to 1.1 with an increment of 0.05. Based on these results, pit by pit discounted cashflow graphs were developed, incorporating a 5% discount rate and a 3% royalty, which allowed the development of potential cash flows for each pit shell iteration.

Once the pits had been generated, and after simulating several planning scenarios for each of them, the extraction rate that would allow for optimal operation of most of the pits was evaluated.

The maximum mining rate, including mineralized material and waste rock, was set at 167,000 tonnes per day (t/d), or 60 Mt per year, to allow for maximal theoretical production rate per pit. However, that mining rate was never reached as the 8000 t/d processing plant throughput was the bottleneck in all scenarios.

16.5.2.1Whittle North Pit Final Shell

The Whittle pit optimization of the northern pit shows a significant increase in waste mined (26.7 Mt) between Pits 14 and 15 relative to the additional mineralized material accessed (0.88 Mt). The graph Figure 16-3 shows a turning point at Pit 14.

Operating cash flow for pit 14 is $141.6M with a RF of 0.9. The pit’s life is approximately one year at full production. There are no opportunities for a pushback. The average stripping ratio for the selected pit shell is 15:1.

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Figure 16-3    North Pit Whittle Shells Iteration

Figure 16-4    North Pit Whittle Shell 14

16.5.2.2Whittle Central Pit Final Shell

The Central Pit, Figure 16-5 shows that pit 15 and pit 16 are the turning point for best case and worst-case scenarios. Provided an optimal pushback sequence, Pit 16 would bring in more revenue, and hence was selected

Operating cash flow for pit 16 is $1.7B with a RF of 1.0. The pit’s mine life is approximately 7 years at full production, with a stripping ratio of 24:1.

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Figure 16-5    Central Pit Whittle Shells Iteration

Figure 16-6    Central Pit Whittle Shell 16

16.5.2.3Whittle South Pit Final Shell

Figure 16-7 illustrates the Whittle iterations for the Southern Pit. The graph shows that beyond pit shell 11 the incremental gain in operating cashflow is marginal. The revenue difference between pit 11 and pit 16 is only $6M, for an added 22Mt of waste.

Operating cash flow for pit 11 is $424M with a RF of 0.75. The pit’s mine life is approximately 1.7 years at full production. The stripping ratio will be 30:1.

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Figure 16-7    South Pit Whittle Shells Iteration

Figure 16-8    South Pit Whittle Shell 11

16.5.3Open Pit Mine Design Parameters

Pit designs were developed for the three deposit areas and are based on the optimized shells and the geotechnical design parameters presented in Table 16-6.

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Table 16-6    Open Pit Design Parameters

Parameter			Value
Layback in overburden			2H:1V
Bench height in overburden			10 m
Bench width between the overburden and rock			Equal to overburden width
Bench in rock			Double 10 m (20 m)
Face angle			70° – 80°
Minimum catchment width			0.2 x Bench height + 4.5m
Inter-ramp slope angle			< 48° for east dipping slopes
Slope break			Every 100 m
Slope break catchment width			Double width of a standard catchment
Pre-split blasting			None
Ramp gradient			10%
Turning radius			20 m
Ramp width double lane (90 t size trucks)			20 m
Ramp width single lane			12 m
Minimum operational bench width			30 m

The mine design process is iterative and aims to transform the optimal pit shells into an operational open pit project. Once completed, the total tonnage of the designed pit should not differ significantly from that of the optimized pit shells. The detailed pit designs were carried out using Deswik mining software. It includes haulage ramps providing access to all benches, except for the final benches which will be excavated via temporary ramps.

16.5.4Haul Road Design

The haulage ramps are sized to accommodate the largest haulage truck (xx t capacity) and are designed for two-way traffic, except for the final levels of each phase, which are designed for one-way traffic. For the two-way traffic, the design considers a buffer space equivalent to half the width of the largest truck width (CAT 785). Safety berms are designed at half the diameter of the haul truck tires. Figure 16-9 (a) and (b) show a schematic of the haul ways for double and single lane.

(a)      (b)

Figure 16-9    Haul Road Schematic (a) and (b)

16.5.5Open Pit Design

Figure 16-10 represents the final mine designs in plan view for North, South and Central pits. The North pit is split into 3 sub-pits each mined in a single phase, the Center pit is mined in 4 phases, and the South pit is mined in two phases. A sample cross section of the Center pit’s phases is shown in Figure 16-11.

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Figure 16-10    Boumadine Pit Locations

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Figure 16-11    Boumadine Pit Center Phases – Longitudinal Section

16.5.6Open Pit equipment Selection

At this stage of the project, it is anticipated that the pits will be mined by contractors. For design purposes, it is assumed that a 200 tonnes CAT 6020 shovel will be paired with a 90 tonne CAT 777 Truck, or similar equipment, will be used. The final choice will depend on the contractor equipment selection to sustain the nominal mill capacity at 8,000 tonnes per day of mineralized material.

16.6Underground Mining

The currently identified underground portion of the Boumadine deposit extends to approximately 640 m below the surface and extends for approximately 5.6 km along strike. It is divided into three main zones: North, Central, and South. The dip of the mineralized zone is relatively steep, ranging from approximately 70° to approximately 80°, and its average width is between approximately 1 and 4 m. The highest-grade mineralized zones are concentrated in the Southern and Central zones, beneath the pits, as shown in Figure 16-12.

Figure 16‑12    Boumadine Underground Mine Design – Longitudinal View

16.6.1Underground Mining Method Description

Given the geometry of the deposit (dip and width) and its geotechnical parameters (rock mass characterization), the long-hole stoping method was deemed appropriate. Because of its productivity and flexibility, the modified Avoca method was selected for this PEA.

This method divides the deposit into sublevels, and the mineralized material between them is mined as stopes. The mineralized material on each sublevel will be mined with longitudinal stopes extending along the strike of the deposit. Each series of longitudinal stopes requires a production drift on the upper sublevel for drilling, loading explosives, and backfilling, and a production drift on the lower sublevel for mucking blasted mineralized material. These production drifts will be driven from a central access drift connected to the ramp.

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Figure 16‑13    Modified Avoca Mining Method Illustration – Longitudinal Section

As shown in Figure 16-13, the longitudinal long-hole stopes will be mined in a retreating fashion towards the access drift. Within each mining panel a bottom-up stoping sequence will be used, with mining commencing on the lowest level of the panel and progressing upward, offset horizontally by three stopes. To initiate each level, a slot raise is drilled and blasted to form a void; subsequence blast hole rings are fired until planned length is achieved based on the rock mass condition of the footwall (FW) and hanging wall (HW). If the ground is competent, stope will be filled entirely of rockfill, the subsequent stope drilled and loaded, and a portion of the rockfill extracted to create a void for blasting. When there is excessive dilution in a panel, the stope will be filled with approximately 50% rockfill (RF) and 50% cemented rockfill (CRF) to offer more support.

The mining cycle of drilling, blasting, mucking and backfilling repeats until the entire level is mined out. The last level will be mined by drilling the blastholes upwards, removing the mineralized material beneath the sill that would have been backfilled with 100% cemented rock fill (CRF). Figure 16-14 illustrates the stope cycle for the proposed mining method.

Figure 16‑14    Modified Avoca Mining Method – Stope Cycle

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This type of mining method requires a regular shape of stopes and mineralized material boundaries. When using mechanical equipment, the drifts need to be large enough to accommodate the required equipment; which can result in dilution where the vein is narrow.

Until further geotechnical recommendations are provided, it is assumed that the ground is overall competent.

16.6.2Underground Stope Design

Stope dimensions were defined based on the recommendations of the geomechanics’ report from RockEng. Each stope was designed to be 20 m high and 20 m long.

Stopes shapes were generated using Deswik Stope Optimizer (DSO) software, taking into account the geometry of the deposit as well as geotechnical and economic criteria.

Several DSO scenarios were evaluated to determine the optimal results in terms of tonnage and grade.

The shapes generated by DSO are visually verified and sequenced to align with the mining method (long-hole longitudinal retreat), then integrated into Deswik Scheduler to establish the production profile and mine life. Table 16-7 summarizes the stope design parameters.

Table 16-7    Stope Design Parameters

UG Stope Design parameters
Parameter			Value
Stope parameters
Mining Method			Modified Avoca
UG NSR COG			124.78 US$/t
Minimum Width			2 m
Maximum Width			20 m
Stope Height (Floor to Floor)			20 m
Stope Length			20 m
Pillar between veins			8 m
Minimum dip angle			54 deg
Mining Dilution ELOS HW			0.5 m
Mining Dilution ELOS FW			0.0 m
Mining recovery			95%
Backfill Dilution			5%

More than 3500 stopes generated were kept in the current mine design. Table 16-8 provides a summary of the shape properties for the different zones.

Table 16-8    Deswik Stope Optimizer (DSO) Output by Zones D

			DSO Generated - Kept			DSO Shape Width (m)									DSO Shape Dip (deg)
			Unit			Min			Max			Average			Min			Max			Average
Centre			770			2.5			8.5			3.8			54			86			65
North			1235			2.5			11.1			3.6			54			87			77
South			1588			2.5			15.5			3.9			54			87			79

Prior to conducting a preliminary economic stope analysis additional stopes were generated between $110 and $124.78 (UG COG), allowing for the evaluation of marginally economic stopes. Isolated stopes and stopes within open pit designs were removed from the underground design. A preliminary economic analysis was then conducted on all the stopes by region, by level and case-by-case to determine if sufficient value is generated to cover the development and infrastructure costs. As a result of this analysis, a few unprofitable

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panels were removed from the production plan. Marginal stopes were then evaluated, after considering the sunk cost of development for stopes above COG, and those that passed the economic evaluation were kept. The mine design was then created for all the remaining stopes.

16.6.3Underground Mine Development Design Parameters

Lateral development will include the main access ramps (declines), level access drifts, the production drifts, and all the necessary excavations for infrastructures.

Vertical development will consist of ventilation raises, which are planned to be excavated using either raise boring, or conventional drill and blast raise (drop raise).

All lateral and vertical development activities are to be performed by a contractor, with the exception of sills, which will be developed using the owner’s teams. Aya provided the unit rates for these activities, based on quotes received from selected contractors, and their experience at their Zgounder mine.

Decline, with a maximum gradient of -15%, will provide access to the mining levels for all zones. The mining area will be accessed by driving level access and sill drifts above and below the stoping area.

For the purposes of this PEA, preliminary ground support design was developed based on the recommendations of the geotechnical report. Standard support in the mining drifts consists of coated Swellex bolts installed in a 1.25 m x 1.25 m square pattern, with mesh applied to the back and walls down to half the height of the drift.

At intersections and in areas with wide spans, additional support will be provided with cable bolts. The recommended length of cables bolts will be half the span on a pattern of 2.0 m × 2.0 m.

The dimensions of the development headings are shown in Table 16-9.

Table 16-9    Development Headings Profiles

			Tunnel Profile (without overbreak)
			Width (m)			Height (m)			Length (m)			Max grade (%)
Lateral Development
Spiral Ramp			4.5			4			Variable			±15
Level Access			4.5			4			Variable			±3
Ventilation Drift			4.5			4			Variable			±3
Connection Drift			4.5			4			Variable			±15
Sill Drift			4			4			Variable			+2
Electrical Main Substation			4.5			4			10			+3
Main Pumping Station			4.5			4			20			-17
Remuck Bay			4.5			4			15			±3
Level Sump (Gravity Drain/Borehole)			4.5			4			15			-15
Temporary Electrical Station			4.5			4			10			±3
Portable Refuge Bay			4.5			4			7			±3
Safety Bay			1.5			2			1			±0
Vertical Development
Ventilation Raises			3 and 4 m Diameter Raise Bore, 3x3 or 3.5x3.5 or 4x4m Drop Raises									NA

Each level includes level access, dewatering sump, ventilation accesses, truck turnaround, remuck bays, and sill drive. Electrical stations will be strategically located so that they could provide power to multiple levels or sub-levels. Figure 16-15 presents a typical mining level.

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Figure 16‑15     Typical Level Layout

16.6.4Emergency Egress

Production activities will respect the current legislation by ensuring that at least two distinct egresses are always available. Thus, to start production as early as possible in some mining areas, some of the internal ventilation raises will need to be outfitted with manways. This will add flexibility to the production sequence while also multiplying egress routes for additional safety. The main access ramp and the main ventilation raise will be considered as the two emergency egress routes out of the mine.

In addition to the egress provisions described above, the following mine safety infrastructure is included in the design:

•Ventilation raises connecting the sublevels will be equipped with ladderways, providing an emergency escape alternative to the spiral ramp.

•Mine rescue equipment will be provided on site for use by the mine rescue team.

•A dedicated ambulance will be stationed at the mine site for emergency medical response.

•Portable refuge stations will be positioned throughout the mine, each designed and equipped to accommodate up to eight (8) people for 36 hours.

•A stench gas system will be installed to provide warning underground in an emergency.

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16.6.5Additional Infrastructures

Additional infrastructure includes emergency underground refuge stations, and internal ventilation raises. Each underground refuge station is designed and located to accommodate the necessary number of workers at any given time. The refuge stations are located no further than a 1,000 m from any location. Powder & cap magazines are not included in the design, but they are considered in the equivalent-metre contingency, and their costs are included in the cash flow model.

16.6.6Dewatering

Mine dewatering system at Boumadine is planned with typical pumping station and submersible pumps located on each level. Dirty water from each level will be gravity-drained to a main sump located at different elevation in each zone. The solids will be decanted and the water pumped to the surface for final effluent treatment or recycled underground for mining operations.

As required, additional small sumps are added, collecting water through ditches by gravity at each low point. The liquid level inside the pumping station sump is monitored and controlled with level transmitters, indicators and controllers, so the submersible pumps can start and stop accordingly. Figure 16-16 illustrates a schematic of the underground mine dewatering for the Center zone. In the Northern and Southern zones, the dewatering process is essentially the same as the central zone. Table 16-10 shows the number of pumping stations for each zone.

Figure 16‑16     Dewatering Network Schematic Central Zone

Table 16‑10    Mine Dewatering

UG Mine			# of pumping stations
UG Center			18
UG South			32
UG North			15

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16.6.7Stope Backfilling

Considering that the mining method adopted for the underground mine is modified Avoca method, the type of backfill will be rock fill (RF) and cemented rock fill (CRF), using waste rock from either the underground mine development or the OP operations.

16.6.7.1Cement Slurry Process

CRF slurry will be mixed in a surface batch plant and then stored in an agitated tank for delivery into a transmixer. The transmixer will transport the CRF slurry UG and dump it onto a measured amount of waste rock stored in a mixing bay near the stope to be backfilled. It is important that the correct proportion of CRF slurry be added to the correct amount of waste rock to ensure the cured target strength of the CRF is achieved. An LHD will thoroughly mix the CRF slurry with the waste rock, ensuring a homogeneous mixture is delivered to the stope. Figure 16-18 represents a schematic of the typical placement of cemented rock fill. Figure 16-17 shows the process of batching and transporting the cement slurry underground.

Figure 16‑17    Cement Slurry Process and Transportation Underground

16.6.7.2Underground CRF Mixing Sump

As illustrated in Figure 16-18, the mix of waste and cement slurry is made in a mixing sump. This process offers the lowest cost form of exposable backfill available for small narrow vein stoping operations. Due to

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the variable nature of run-of-mine and limited control over the sump mixing process, CRF is generally not suitable for underhand exposure in man-entry stopes. The variability of the CRF material is limited by using only good quality waste rock that has no clay component and screening particles greater than 400 mm from the waste. Ongoing quality control and assurance is critical to the safe and efficient implementation of CRF. A thorough QA/QC procedure must be developed that includes inspection of waste material, sampling of the cement slurry, reconciliation of cement consumption, sampling and laboratory testing of CRF material, and visual inspection of exposure performance.

Figure 16‑18     Mixing Sump Schematic

16.6.7.3Stope Backfill Placement

Rock fill will be placed in the stopes by direct dumping from access levels, depending on stope geometry and accessibility. For cemented rockfill, the material will consist of cemented waste rock and waste rock. Placement will generally proceed in lifts, with each lift allowed to settle before subsequent filling to maintain stability and prevent backfill dilution. The cemented rock fill and rock fill mass will act primarily as passive support rather than structural strength during adjacent stope extraction. The use of rock fill as a primary backfilling option will improve the mining cycle by reducing the quantity of waste that must be rehandled. Figure 16-19 shows a schematic of the ratio of CRF and RF that will be used in poor ground conditions.

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Figure 16‑19    Stope Backfill Placement in Poor Ground Conditions

As mentioned earlier, the first level will be fully cemented to permit the recovery of the sill pillar below the mining horizon. CCE recommends evaluating the use of surface backfill raises to ensure a performant and efficient underground waste rock distribution. The overall system must provide operational flexibility and efficiency enabling continuous backfilling concurrent with extraction.

16.6.8Underground Mine Equipment

The underground mobile equipment estimate has been detailed for all activities. A portion of the fleet will be owned and operated by the development contractors. Raiseboring development will be completed by contractors. The Mobile equipment units needed per year for each type of equipment are based on their estimated productivity and the planned production targets outlined in the LOM plan. Additional units are included in the cost estimate to account for availability factors. The LOM mobile equipment fleet is summarized in Table 16‑11.

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Table 16‑11    Operating mobile equipment units – Annual requirements

Description			Model			2029			2030			2031			2032			2033			2034			2035			2036			2037			2038			2039			2040			2041			2042			2043
Jumbo			Sandvik DD321			-			-			2			4			5			5			6			7			7			8			8			8			8			6			1
Production Drill			Sandvik DL311			-			-			-			1			2			2			2			3			3			3			3			3			3			3			2
Drill Compressor						-			-			-			1			2			2			2			3			3			3			3			3			3			3			2
Emulsion Charger			PAUS UNI 40ST-LG			-			-			1			2			2			2			2			3			3			3			3			3			3			2			1
33 tonnes Mine Truck			Epiroc MT436B			-			-			1			2			4			5			6			6			8			8			6			6			7			7			2
10 tonnes LHD			Epiroc ST1030			-			-			1			3			5			8			8			9			9			10			10			10			10			10			4
Bolter			Sandvik DS 421			-			-			3			5			6			6			7			8			8			9			9			9			8			7			2
Transmixer			Partindus HURON 45			-			-			1			1			1			1			1			1			1			1			1			1			1			1			1
Scissor Lift			PAUS Scissor Lift			-			-			2			2			3			3			3			3			3			3			3			3			3			2			1
Boom Truck			PAUS TSL 853 T7			-			-			5			5			5			5			5			5			5			5			5			5			5			5			5
Grader			XCMG GR135C			-			-			1			1			1			1			1			1			1			1			1			1			1			1			1
Fuel Lube Truck			PAUS UNI 40SV			-			-			1			1			1			1			1			1			1			1			1			1			1			1			1
Water Truck			PAUS UNI 40SV			-			-			1			1			1			1			1			1			1			1			1			1			1			1			1
Pickup Truck			Toyota Hilux			-			-			15			15			15			15			15			15			15			15			15			15			15			15			15
Mine rescue Truck			ACCESS			-			-			1			1			1			1			1			1			1			1			1			1			1			1			1
Total						-			-			35			45			54			58			61			67			69			72			70			70			70			65			40

Note: It is assumed that the raiseboring development will be completed by a contractor.

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16.6.9Underground Mine Personnel

The mine will operate year-round, seven (7) days per week, twenty-four (24) hours per day, two (2) shifts of 12-hour shifts. Most of the staff will be recruited locally, supplemented by a few expatriate personnel were needed. In addition, Aya plans to mobilize experienced personnel from its Zgounder mine in Morocco to provide on-site training and help ensure that operations are conducted safely and efficiently.

Table 16-12 shows the staffing requirements for full production, covering underground production as well as technical, supervisory, and support staff. It excludes contractor personnel involved in surface production and underground development. This table considers absences, leave, and temporary vacancies.

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Table 16‑12    Maximum Labour Requirements

Task			2029			2030			2031			2032			2033			2034			2035			2036			2037			2038			2039			2040			2041			2042			2043
Open Pit Employees Owner Only
O/P Superintendent			1			1			1			1			1			1			1			1			1			1			1			-			-			-			-
O/P Mine Administrative			1			1			1			1			1			1			1			1			1			1			1			-			-			-			-
O/P Captain			1			1			1			1			1			1			1			1			1			1			1			-			-			-			-
O/P Technical Services			10			10			10			10			10			10			10			10			10			10			10			-			-			-			-
Subtotal			13			13			13			13			13			13			13			13			13			13			13			-			-			-			-
Underground Employees Owner Only
Staff
U/G Superintendent			-			1			1			1			1			1			1			1			1			1			1			1			1			1			1
U/G Mine Supervisor			-			1			9			14			16			20			22			23			25			25			25			25			24			24			13
U/G Technical Services			-			-			-			36			36			36			36			36			36			36			36			36			36			36			36
Maintenance Supervisor			-			1			2			5			6			6			6			6			6			6			6			6			6			6			6
Maintenance Planner			-			-			-			2			2			2			2			2			2			2			2			2			2			2			2
Subtotal			-			3			12			58			61			65			67			68			70			70			70			70			69			69			58
Mine Operating
Dev. Jumbo			-			-			4			8			8			10			13			16			16			15			15			15			15			12			-
Raisebore			-			-			-			-			-
Production Drill			-			-			-			3			4			8			8			10			10			12			11			10			11			12			5
Production Drill Services			-			-			-			3			4			8			8			10			10			12			11			10			11			12			5
Emulsion Loading			-			-			-			3			3			5			5			6			7			7			7			7			7			6			1
Tramming (33t Truck)			-			-			2			8			13			20			23			23			31			32			26			26			26			26			8
LHD (10t)			-			-			1			9			15			25			27			30			33			36			33			33			33			33			15
Bolter			-			-			4			12			12			14			17			20			20			24			24			24			19			16			4
Transmixer			-			-			4			4			4			4			4			4			4			4			4			4			4			4			4
Scissor Lift			-			-			4			6			6			7			7			8			8			8			8			8			8			7			4
Boom Truck			-			-			40			40			40			40			40			40			40			40			40			40			40			40			40
Grader			-			-			8			8			8			8			8			8			8			8			8			8			8			8			8
Fuel/Lube Truck			-			-			4			4			4			4			4			4			4			4			4			4			4			4			4
Water Truck			-			-			4			4			4			4			4			4			4			4			4			4			4			4			4
Subtotal			-			-			75			112			125			157			168			183			195			206			195			193			190			184			102
Maintenance
Mobile Mechanic			-			1			9			11			13			15			16			17			18			18			18			18			18			18			11
Fixed Mechanic			-			-			-			8			8			8			8			8			8			8			8			8			8			8			8
Fixed Mech. Batch Plant			-			-			-			4			4			4			4			4			4			4			4			4			4			4			4
Mobile Electrician			-			1			3			4			5			5			6			6			6			6			6			6			6			6			4
Fixed Electrician			-			-			-			8			8			8			8			8			8			8			8			8			8			8			8
Fixed Elec. Batch Plant			-			-			-			4			4			4			4			4			4			4			4			4			4			4			4
Subtotal			-			2			12			39			42			44			46			47			48			48			48			48			48			48			39
Total AYA O/P & U/G			13			18			112			222			241			279			295			312			326			337			326			311			307			301			199

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16.7Stockpile Management

Net smelter return (NSR) was used to assess the economic potential of the Boumadine project. The NSR value represents the expected net value of the payable metal contained the in saleable concentrates, after deducting treatment, refining and penalty charges.

The NSR-Mine was calculated as an additional block model field, allowing a single value representing lead, zinc, gold and silver grades to be used for mine planning, scheduling, and stockpile classification. A second value, the NSR-Pyrite, was calculated to further classify the higher-value material based on the value of its pyrite concentrate. This section describes the calculation of the NSR-Mine and NSR-Pyrite, followed by the stockpile classification criteria.

16.7.1NSR Calculation – Production Plan

The NSR-Mine formula was derived by estimating the net revenue generated per unit of head grade for each payable metal, based on the three saleable concentrates expected from the process plant: a lead concentrate (payable lead, gold and silver), a zinc concentrate (payable zinc) and a pyrite concentrate (payable gold and silver). The metal prices, refining charges and treatment charges applied to each concentrate are presented in Table 16-13.

Table 16-13    Metal Prices and Smelter Terms by Concentrate

Element			Metal Price			Refining Charge			Price net of RC
Lead Concentrate – TC US$122/dmt (incl. US$22/dmt penalties); 29.6% Pb concentrate grade
Pb			US$1.00/lb			–			US$1.00/lb
Au			US$2,800/oz			US$30.00/oz			US$2,770/oz
Ag			US$30.00/oz			US$1.50/oz			US$28.50/oz
Zinc Concentrate – TC US$205.50/dmt (incl. US$7/dmt penalties); 57.4% Zn concentrate grade
Zn			US$1.20/lb			–			US$1.20/lb
Pyrite Concentrate – no treatment or refining charges
Au			US$2,800/oz			–			US$2,800/oz
Ag			US$30.00/oz			–			US$30.00/oz

Gross payable values were calculated from the metallurgical recovery, smelter payable and metal price net of refining charges, converted to the appropriate grade unit (22.046 lb per 1% of a tonne for lead and zinc, and 31.1035 g per troy ounce for gold and silver). Treatment charges are levied per dry tonne of concentrate and, at a constant recovery and concentrate grade, are directly proportional to the lead or zinc head grade. They were therefore expressed per percent of head grade and deducted from the lead and zinc factors, respectively. No treatment or refining charges were applied to the pyrite concentrate.

Concentrate transportation, storage, ship loading and ocean freight costs were set to zero in the NSR calculation.

The derivation of each NSR-Mine factor is summarized in Table 16-14.

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Table 16-14    NSR-Mine Factor Derivation

Concentrate			Recovery (%)			Payable (%)			Price net of RC			Gross Value			Less TC			NSR Factor
Lead – US$ per % Pb
Lead			82.0			90.0			US$1.00/lb			16.25			(3.38)			12.87
Zinc – US$ per % Zn
Zinc			74.7			85.0			US$1.20/lb			16.80			(2.67)			14.12
Gold – US$ per g/t Au
Lead			23.8			95.0			US$2,770/oz			20.14			–			20.14
Pyrite			71.4			78.0			US$2,800/oz			50.14			–			50.14
Total												70.27			–			70.27
Silver – US$ per g/t Ag
Lead			50.4			95.0			US$28.50/oz			0.44			–			0.44
Pyrite			41.4			78.0			US$30.00/oz			0.31			–			0.31
Total												0.75			–			0.75

Note: Treatment charge deductions are based on the treatment charges and concentrate grades in Table 16-13. Totals may not add due to rounding.

The resulting NSR-Mine formula is:

NSR-Mine (US$/t) = (Pb% x $12.87) + (Zn% x $14.12) + (Au g/t x $70.27) + (Ag g/t x $0.75) - $6.75*

* Royalties per tonne of feed.

16.7.2NSR Calculation – Pyrite Concentrate

The NSR-Pyrite is expressed per tonne of pyrite concentrate rather than per tonne of plant feed. It indicates the potential value of the pyrite concentrate generated from each tonne of plant feed. As each tonne of feed produces less than one tonne of pyrite concentrate, the NSR-Pyrite may be higher than the NSR-Mine.

The value of the gold and silver reporting to the pyrite concentrate was first calculated per tonne of feed, using the pyrite concentrate factors from Table 16-14. This value was then divided by the mass of pyrite concentrate produced per tonne of feed, which is estimated from the iron head grade using the iron recovery and pyrite concentrate grade presented in Table 16-15. For example, feed grading 22% Fe produces approximately 0.51 t of pyrite concentrate per tonne.

Table 16-15    NSR-Pyrite Calculation Parameters

Parameter			Unit			Value
Pyrite Concentrate Value (from Table 16-14)
Gold factor			US$ per g/t Au			50.14
Silver factor			US$ per g/t Ag			0.31
Pyrite Concentrate Mass
Fe recovery to pyrite concentrate			%			89.0
Pyrite concentrate grade			% Fe			38.1
Mass factor (89.0% / 38.1% Fe)			t conc./t feed per % Fe			0.02336

The resulting NSR-Pyrite formula is:

NSR-Pyrite (US$/t) = ((Au g/t x $50.14) + (Ag g/t x $0.31)) / (Fe% x 0.02336)

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16.7.3Stockpile Classification

The mineralized material included in the production schedule has been classified into three stockpiles (Low Grade Stockpile, LG Pyrite Stockpile and HG Pyrite Stockpile) based on both the NSR-Mine and the NSR-Pyrite. The classification thresholds for the open pit and underground mines are presented in Table 16-16 and illustrated in Figure 16-20.

Table 16-16    Stockpile Classification Criteria

Stockpile			NSR-Mine			NSR-Pyrite
Open Pit
Low Grade Stockpile			US$60/t ≤ NSR-Mine < US$140/t			–
LG Pyrite Stockpile			NSR-Mine ≥ US$140/t			NSR-Pyrite < US$340/t
HG Pyrite Stockpile			NSR-Mine ≥ US$140/t			NSR-Pyrite ≥ US$340/t
Underground
Low Grade Stockpile			US$77/t ≤ NSR-Mine < US$130/t			–
LG Pyrite Stockpile			NSR-Mine ≥ US$130/t			NSR-Pyrite < US$240/t
HG Pyrite Stockpile			NSR-Mine ≥ US$130/t			NSR-Pyrite ≥ US$240/t

Figure 16‑20    Stockpile Classification Parameters

16.8Production Plan

Both the underground (UG) and open-pit (OP) mining sequences were developed using Deswik Scheduler software. The life-of-mine (LOM) production plan for the Boumadine project extends over fifteen (15) years, including pre-production.

To improve the project overall economics, the production schedule prioritizes the mining of higher-grade mineralized material during the early years. In addition, underground production is intentionally deferred by two (2) years after the start of open-pit operations. This phased approach allows for earlier cash flow

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generation from open-pit mining while delaying the capital-intensive underground development to a later stage of the mine life.

16.8.1Pre-Production Plan

Production is scheduled to begin in early 2029, during the project construction phase (“Year 0” of the financial model). As such, pre-production activities must begin well in advance to allow for the early critical works. The main activities carried out during the pre-production period will include the construction of roads, pads, and surface waste stockpiles. Pre-stripping of the mining areas (OP) will be performed by a contractor. The waste rock dump facilities must be prepared before stripping can begin.

16.8.2Open Pit Production Plan

Open pit mining will begin in two main areas in the Center and South pits, prioritizing higher-grade mineralized material. The Central pit, which contains higher-grade mineralized material, will serve as the main pit, and will be mined first, starting in 2029. Mining in this pit will continue for approximately eleven (11) years. The Central Pit will be mined in four (4) phases.

At the same time, South pit will be developed more gradually to supplement production from Central Pit. The South Pit will be mined in two (2) phases.

South pit will be mined over a period of six (6) years, while mining at North pit will start in 2034. The annual production projected, including both mineralized material and waste is scheduled at a maximum of 55Mt per year.

Figure 16-21 a, b, c, d, e and f present the open pit mining for Central, Southern and Northern pits.

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Figure 16‑21    Open Pit Sequence: Surface view every 2 year: (a) 2030 (b) 2032 (c) 2034 (d) 2036 (e) 2038 (f) 2039

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Table 16-17 shows the breakdown of open-pit life-of-mine (LOM) production by stockpiles, while Table 16-18 shows a production breakdown by pit. All tonnes shown are diluted at 5% with 0% mineralized material loss.

Table 16‑17    Open Pit Life-Of-Mine Production by Stockpile

Description			Units			2029			2030			2031			2032			2033			2034			2035			2036			2037			2038			2039
HG Pyrite Stockpile
Mineralized material tonnage			tonnes			634,432			902,494			511,050			885,797			2,399,064			321,842			658,421			19,820			82,579			804,509			1,367,008
NSR Average			$/tonne			340.4			470.0			262.8			346.7			355.1			273.3			310.4			225.1			208.3			299.0			339.9
LG Pyrite Stockpile
Mineralized material tonnage			tonnes			367,871			1,177,778			667,392			952,227			676,575			1,095,513			958,316			418,073			643,452			1,124,585			384,541
NSR Average			$/tonne			193.4			230.6			191.2			209.9			224.1			192.5			192.0			158.3			174.8			193.0			245.3
Low Grade Stockpile
Mineralized material tonnage			tonnes			330,856			423,248			632,714			437,938			332,672			567,495			759,027			535,048			477,575			656,180			148,003
NSR Average			$/tonne			97.8			97.3			92.5			98.8			93.6			97.9			98.1			100.1			96.8			103.0			109.5
Mineralized material tonnage			tonnes			1,333,158			2,503,519			1,811,155			2,275,963			3,408,312			1,984,850			2,375,764			972,941			1,203,606			2,585,273			1,899,552
NSR Average			$/tonne			239.6			294.4			176.9			241.7			303.6			178.6			194.8			127.7			146.2			203.2			302.8

Table 16‑18    Open Pit Life-Of-Mine Production by Zone

Description			Units			2029			2030			2031			2032			2033			2034			2035			2036			2037			2038			2039			Total
Central Pit
Mineralized material tonnage			tonnes			1,329,316			1,977,788			1,011,004			1,268,656			2,712,296			1,255,810			1,451,152			589,508			622,443			2,585,273			1,899,552			16,702,797
NSR Average			$/tonne			240			324.7			156.3			188.7			286.8			193.3			218.2			127.9			136.7			203.2			302.8			236.9
Waste Mined			tonnes			13,666,926			23,007,125			28,984,998			28,741,745			43,322,528			38,746,713			41,295,737			48,310,678			41,079,349			47,414,706			24,898,476			379,468,981
Total Material			tonnes			14,996,243			24,984,913			29,996,002			30,010,401			46,034,825			40,002,523			42,746,888			48,900,186			41,701,792			49,999,979			26,798,028			396,171,778
South Pit
Mineralized material tonnage			tonnes			3,842			525,731			800,152			1,007,307			696,015			-			-			-			581,163			-			-			3,614,210
NSR Average			$/tonne			104.7			180.1			203			308.6			368.8			-			-			-			156.3			-			-			253.4
Waste Mined			tonnes			5,000,147			19,485,012			24,203,866			23,992,686			8,269,186			-			-			-			12,720,089			-			-			93,670,985
Total Material			tonnes			5,003,989			20,010,743			25,004,017			24,999,993			8,965,201			-			-			-			13,301,252			-			-			97,285,196
North Pit
Mineralized material tonnage			tonnes			-			-			-			-			-			729,041			924,613			383,433			-			-			-			2,037,087
NSR Average			$/tonne			-			-			-			-			-			153.2			158.2			127.2			-			-			-			150.5
Waste Mined			tonnes			-			-			-			-			-			14,270,032			11,328,410			5,718,903			-			-			-			31,317,345
Total Material			tonnes			-			-			-			-			-			14,999,072			12,253,023			6,102,336			-			-			-			33,354,432
Total Tonnage Mined
Mineralized material tonnage			tonnes			1,333,158			2,503,519			1,811,155			2,275,963			3,408,312			1,984,850			2,375,764			972,941			1,203,606			2,585,273			1,899,552			22,354,094
NSR Average			$/tonne			239.6			294.4			176.9			241.7			303.6			178.6			194.8			127.7			146.2			203.2			302.8			231.7
Waste Mined			tonnes			18,667,073			42,492,137			53,188,864			52,734,431			51,591,714			53,016,745			52,624,147			54,029,581			53,799,438			47,414,706			24,898,476			504,457,312
Total Material			tonnes			20,000,232			44,995,656			55,000,019			55,010,394			55,000,025			55,001,596			54,999,911			55,002,522			55,003,044			49,999,979			26,798,028			526,811,406
Stripping Ratio						14			17			29			23			15			27			22			56			45			18			13			23

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Figure 16-22    Mineralized Material Mined per Year and per Pits

Figure 16-23    Waste Mined per Year per Pits

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Figure 16-24    Open Pit Life-of-Mine Production Plan and NSR

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•

16.8.3Underground Development Plan

Development of the underground facility is expected to begin in 2031 and will take approximately one year to reach full production capacity for the first portal. Full ramp up of underground production will end in 2038, as open pit production ramps down. All development activities will be carried out by a contractor, except for sill development. The pre-production phase of underground operations will include the procurement of equipment, the portal excavation and ventilation shafts, as well as the recruitment of qualified mining personnel to support the start-up of underground mining operations. 

Access to all underground mining zones of the Boumadine deposit will be provided by spiral ramps, either from the surface or from existing pits that have already been developed.  

Ramps will provide access to all levels of the mine and, subsequently, to the production sub-levels, which will be spaced 20 meters apart. Also, the ramps will be used for: 

•Hauling mineralized material and waste to surface; 

•Moving equipment and personnel from level to level; 

•Providing return airways and serving as emergency egress;

•Providing a pathway for the mine dewatering. 

Other ancillary infrastructures will be located throughout the mine, such as drainage sumps, main pumping station, electrical substations, powder magazines, lubricant and fuel storage bays.  

Vertical development will consist of a surface ventilation raise, which will be excavated using raise boring machine, and inter level raises that will be developed as drop raises or raise bore depending on the airflow needs and the length of raises. The exhaust air raises will be driven in parallel with the main ramp. In addition, the exhaust raises will connect all levels and sublevels and equipped with ladders to serve as secondary escape routes. 

It is also assumed that a portion the surface waste material will be backhauled underground to supplement the waste generated by development, when required. Figure 16-25 shows the main underground infrastructures for the south zone portal 1. 

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Figure 16-25    Development Heading (South Zone Portal 1)

The lateral development rate used for the Boumadine project is based on the Zgounder mine. Table 16-19 presents the performance used in Deswik Mine Scheduler. 

Table 16-19    Development Performance Rate 

Activity / Per Dev Team			Unit			Underground Rate
Decline and Mine Access			m/day			8
Lateral waste development, multiple headings			m/day			8
Sill development, multiple headings			m/day			8
Vertical Development			m/day			3

The dilution factor and the recovery for the mineralized development used are at 0% and 100% respectively, and the incremental NSR cut-off applied for development is 77 $/tonne. In addition, an uplift factor was used to account for added meter from infrastructure that was not designed at this stage and additional Opex meters. A factor of 10% was used on the Capex meters and 5% on the Opex meters.

Table 16-20 presents the LOM development plan summarized by development type and categorized into Capital and Operating. 

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Table 16‑20    Life-of-Mine Underground Development Plan

Description			Unit			2029			2030			2031			2032			2033			2034			2035			2036			2037			2038			2039			2040			2041			2042			2043			Total
Lateral Development
Ramp			m			-			-			3,563			3,578			5,354			4,394			3,651			4,686			5,736			5,716			5,480			5,983			7,076			3,065			-			58,282
Level Access			m			-			-			727			1,320			1,870			1,347			3,513			2,615			1,655			3,350			4,311			2,481			1,999			1,619			-			26,807
Sill			m			-			-			-			5,819			5,840			7,496			9,496			11,712			11,680			10,753			10,975			14,307			11,177			9,030			322			108,607
Ventilation Drift			m			-			-			597			465			843			1,046			989			524			874			1,095			1,280			2,582			2,164			1,387			-			13,846
Dev. Infrastructure			m			-			-			313			493			693			368			582			614			495			895			610			612			441			513			-			6,629
Total Lateral Dev.			m			-			-			5,200			11,675			14,600			14,651			18,231			20,151			20,440			21,809			22,656			25,965			22,857			15,614			322			214,171
Vertical Development
Ventilation Raises			m			-			-			266			433			945			875			464			1,092			926			1,286			356			1,091			1,319			914			-			9,967
Dewatering Holes			m			-			-			173			292			528			160			236			372			242			541			218			211			-			270			-			3,243
CAPEX vs OPEX Development
CAPEX Dev.			m			-			-			5,639			6,581			10,233			8,190			9,435			9,903			9,928			12,883			12,255			12,960			12,999			7,768			-			118,774
OPEX Dev.			m			-			-			-			5,819			5,840			7,496			9,496			11,712			11,680			10,753			10,975			14,307			11,177			9,030			322			108,607

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16.8.4Underground Production Plan

Since the underground mine will operate in parallel with open-pit mining activities, zone access ramps will start construction from the surface. This approach ensures rapid access to the underground mining areas without being constrained by the sequence of open-pit mining operations. There will be five portals providing surface access: Portals 1 to 3 in the South Zone, Portal 4 in the Center Zone, and Portal 5 in the North Zone. The zones and their respective portal numbers can be seen in Figures 16-26, 16-27, and 16-28 below.

The underground mining sequence will follow the same strategy as the open-pit operation, prioritizing high-grade zones. The sequence of each portal was independently evaluated and optimized prior to integration in the overall sequence. When considering only average NSR, the ideal sequence would be Portal 1, 3, 2, 4, and then 5. However, additional considerations such as accessibility to higher and lower grade areas, resource utilization, development requirements and timelines, and yearly consistency in production in terms of grade and tonnage was included in the evaluation of the optimized underground sequence. As such, underground operations will begin in Portal 1 in 2031, with mineralized material production in 2032. Portal 1 is followed by Portal 4 in 2034, Portal 3 in 2036, Portal 2 in late 2036, and Portal 5 in 2038. Portals 4 and 5 will continue operating in tandem until underground mining operations are complete in 2043.

Figure 16‑26    Primary UG Mining Area in the South Zone

In the early years of the schedule, as mineralized material production and capital spending are ramping up, Portal 1 will be the sole focus of mining operations. Portal 4 will begin development in 2034 to achieve maximum mineralized material production in 2036 when Portal 1’s production capabilities start diminishing. Operations in Portal 3 will commence in 2036, followed by Portal 2 towards the end of the year as more equipment resources are freed from Portal 1 in the lead up to its completion in early 2037. Priority shifts to Portals 3 and 2 in 2037, throttling Portal 4’s production as higher-grade mineralized material becomes depleted in the Center Zone.

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Figure 16‑27    Secondary UG Mining Area in the Center Zone

Portal 5 will begin development in 2038 and start ramping up its mineralized material contribution as mining operations in the South Zone come to completion with Portal 3 in 2039, and Portal 2 in 2040. After which Portal 5 remains the priority, focusing on its higher-grade areas and finishing together with the remainder of Portal 4 in 2043.

Figure 16‑28    Remaining UG Mining Area in the North Zone

Where permitted and favorable to the mine sequence, deep mineralized zones have been divided into two or more mining horizons, separated by a sill pillar. The first mineable level of the mining horizon will be backfilled

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with Cemented Rock Fill (CRF), allowing for efficient and safe mining in the levels below. This strategy is important to reduce development requirements prior to mineralized material production in vertically long mineralized zones while enabling earlier access to high-grade areas.

The stope production cycle was estimated based on a typical stope with an average width of 3.6m, a height of 16 m (from the floor to ceiling), and a length of 20 m. The stope parameters were based on an average density of the mineralized material of 3.7 t/m³, a dilution of 5%, a recovery rate of 95%, and a swell factor of 1.365. Table 16-21 presents the production cycle parameters used in Deswik Scheduler. 

Table 16‑21    Stope Production Cycle Parameters

Activity			Unit			Rate			Duration (day)
Stope Preparation			d/stope			2			2
Production Drilling			m/d/drill			345			2
Blasting			d/stope			0.6			0.6
Mucking			t/d			1530			2.8
Backfill Preparation			d/stope			1			1
Average Stope Backfill			t/d			1 637			1.6
Stope Curing Time			Day			14			14
Stope Cycle (average)			Day						24

Table 16-22 and 16-23 present the underground production profile by stockpiles and by zones. 

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Table 16‑22    Underground Production Plan Sorted by Stockpile Category

Description			Unit			2029			2030			2031			2032			2033			2034			2035			2036			2037			2038			2039			2040			2041			2042			2043			Total
HG Pyrite Stockpile
Mineralized material tonnage			tonnes			-			-			-			166,212			219,782			160,924			295,852			698,010			615,182			656,802			230,231			252,717			146,284			231,213			16,655			3,689,864
NSR Average			$/tonne			-			-			-			237.4			245.3			189.8			172.3			186.0			203.5			194.8			177.5			191.2			160.2			167.9			133.6			192.8
LG Pyrite Stockpile
Mineralized material tonnage			tonnes			-			-			-			300,330			372,721			946,072			726,048			550,009			517,641			641,570			586,623			892,347			1,097,850			650,140			415,637			7,696,988
NSR Average			$/tonne			-			-			-			185.2			164.7			160.8			150.7			157.8			161.7			154.0			148.7			160.0			162.6			150.3			145.5			157.8
Low Grade Stockpile
Mineralized material tonnage			tonnes			-			-			-			100,416			232,578			364,892			496,395			626,740			712,287			701,525			1,119,196			854,776			755,672			1,118,450			424,942			7,507,869
NSR Average			$/tonne			-			-			-			109.9			113.9			115.8			113.1			113.0			115.1			116.3			114.1			111.6			113.5			114.8			117.1			114.2
Total Mineralized Material Mined
Mineralized material tonnage			tonnes			-			-			-			566,958			825,081			1,471,888			1,518,295			1,874,759			1,845,110			1,999,897			1,936,050			1,999,840			1,999,806			1,999,803			857,234			18,894,721
NSR Average			$/tonne			-			-			-			187.2			171.9			152.8			142.6			153.3			157.6			154.2			132.1			143.3			143.9			132.5			131.2			147.3

Table 16‑23    Underground Production Plan Sorted by Mining Zone

Description			Unit			2029			2030			2031			2032			2033			2034			2035			2036			2037			2038			2039			2040			2041			2042			2043			Total
Central Zone
Mineralized material tonnage			tonnes			-			-			-			-			-			-			96,430			1,276,745			1,268,514			547,656			702,639			-			-			208,532			154,242			4,254,758
NSR Average			$/tonne			-			-			-			-			-			-			109.2			155.8			157.9			138.8			126.4			-			-			119.7			141.5			146.0
Waste Mined			tonnes			-			-			-			-			-			164,120			431,639			215,244			57,432			43,202			3,557			92,324			3,076			227,211			2,639			1,240,444
Total Material Moved			tonnes			-			-			-			-			-			164,120			528,069			1,491,989			1,325,946			590,858			706,196			92,324			3,076			435,743			156,881			5,495,202
South Zone
Mineralized material tonnage			tonnes			-			-			-			566,958			825,081			1,471,888			1,421,885			598,014			576,597			1,451,935			821,303			668,576			-			-			-			8,402,237
NSR Average			$/tonne			-			-			-			187.2			171.9			152.9			144.9			148.1			157.1			159.9			138.9			148.9			-			-			-			155.2
Waste Mined			tonnes			-			-			248,590			334,713			495,505			246,921			82,996			289,428			547,596			408,009			46,203			25,945			-			-			-			2,725,906
Total Material Moved			tonnes			-			-			248,590			901,671			1,320,586			1,718,809			1,504,881			887,442			1,124,193			1,859,944			867,506			694,521			-			-			-			11,128,143
North Zone
Mineralized material tonnage			tonnes			-			-			-			-			-			-			-			-			-			306			412,108			1,331,265			1,999,806			1,791,271			702,992			6,237,748
NSR Average			$/tonne			-			-			-			-			-			-			-			-			-			168.4			128.3			140.4			143.9			133.9			128.9			137.6
Waste Mined			tonnes			-			-			-			-			-			-			-			-			-			262,807			613,500			571,408			687,164			269,156			2,782			2,406,817
Total Material Moved			tonnes			-			-			-			-			-			-			-			-			-			263,113			1,025,608			1,902,673			2,686,970			2,060,427			705,774			8,644,565
Total Mineralized Material Mined
Mineralized material tonnage			tonnes			-			-			-			566,958			825,081			1,471,888			1,518,315			1,874,759			1,845,111			1,999,897			1,936,050			1,999,841			1,999,806			1,999,803			857,234			18,894,743
NSR Average			$/tonne			-			-			-			187.2			171.9			152.9			142.6			153.3			157.7			154.1			132.1			143.2			143.9			132.4			131.2			147.3
Waste Mined			tonnes			-			-			248,590			334,713			495,505			411,041			514,635			504,672			605,028			714,018			663,260			689,677			690,240			496,367			5,421			6,373,167
Total Material Moved			tonnes			-			-			248,590			901,671			1,320,586			1,882,929			2,032,950			2,379,431			2,450,139			2,713,915			2,599,310			2,689,518			2,690,046			2,496,170			862,655			25,267,910

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Figure 16-29    Underground Mineralized Material Mined by Zones

Figure 16-30    Underground Mineralized Material Mined and NSR

16.8.5Combined Production Plan

The total mineralized material tonnage for the mine is estimated at 41.2 million tonnes, with an average NSR grade of 193 US$/t. OP mining will contribute with 54.2% of the total tonnage, while UG mining will account for the remaining 45.8%. In the final four years of the mine life, production will be exclusively from UG operations.

Because the tonnages from open pit and underground are not balanced year on year, the strategy is to use the different stockpiles to regulate and optimize the feed to the processing plant, ensuring a consistent tonnage and grade profile over the LOM. Boumadine production plan is summarized in Table 16‑24, Figure 16‑31 and Figure 16‑32.

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Table 16‑24    Boumadine Life-of-Mine Production Plan

Description			Units			2029			2030			2031			2032			2033			2034			2035			2036			2037			2038			2039			2040			2041			2042			2043			Total
Open Pit Mine
Mineralized material tonnage			tonnes			1,333,158			2,503,519			1,811,155			2,275,963			3,408,312			1,984,850			2,375,764			972,941			1,203,606			2,585,273			1,899,552			-			-			-			-			22,354,094
NSR Average			$/tonne			239.6			294.4			176.9			241.7			303.6			178.6			194.8			127.7			146.2			203.2			302.8			-			-			-			-			231.7
Waste Mined			tonnes			18,667,073			42,492,137			53,188,864			52,734,431			51,591,714			53,016,745			52,624,147			54,029,581			53,799,438			47,414,706			24,898,476			-			-			-			-			504,457,312
Total Material Moved			tonnes			20,000,232			44,995,656			55,000,019			55,010,394			55,000,025			55,001,596			54,999,911			55,002,522			55,003,044			49,999,979			26,798,028			-			-			-			-			526,811,406
Underground Mine
Mineralized material tonnage			tonnes			-			-			-			566,958			825,081			1,471,888			1,518,295			1,874,759			1,845,111			1,999,897			1,936,050			1,999,841			1,999,806			1,999,803			857,234			18,894,722
NSR Average			$/tonne			-			-			-			187.2			171.9			152.9			142.6			153.3			157.7			154.1			132.1			143.2			143.9			132.4			131.2			147.3
Waste Mined			tonnes			-			-			248,590			334,713			495,505			411,041			514,635			504,672			605,028			714,018			663,260			689,677			690,240			496,367			5,421			6,373,167
Total Material Moved			tonnes			-			-			248,590			901,671			1,320,586			1,882,929			2,032,930			2,379,431			2,450,139			2,713,915			2,599,310			2,689,518			2,690,046			2,496,170			862,655			25,267,889
Open Pit & Underground Tonnes Mined
Mineralized material tonnage			tonnes			1,333,158			2,503,519			1,811,155			2,842,921			4,233,393			3,456,738			3,894,060			2,847,700			3,048,716			4,585,170			3,835,602			1,999,841			1,999,806			1,999,803			857,234			41,248,816
NSR Average			$/tonne			239.6			294.4			176.9			230.9			277.9			167.6			174.5			144.6			153.1			181.8			216.7			143.2			143.9			132.4			131.2			193.1
Waste Mined			tonnes			18,667,073			42,492,137			53,437,454			53,069,144			52,087,219			53,427,786			53,138,782			54,534,253			54,404,466			48,128,724			25,561,736			689,677			690,240			496,367			5,421			510,830,479
Total Material Moved			tonnes			20,000,232			44,995,656			55,248,609			55,912,065			56,320,612			56,884,524			57,032,842			57,381,953			57,453,183			52,713,894			29,397,338			2,689,518			2,690,046			2,496,170			862,655			552,079,295

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Figure 16‑31    Boumadine Life-Of-Mine Production Plan Breakdown by Mining Methods

Figure 16‑32    Boumadine LOM Production Plan Tonnes and NSR

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16.9Electrical Underground Distribution

The electrical infrastructure for the Boumadine mine has been designed around a modular, relocatable concept that follows the development sequence. Power will be taken from the surface 22 kV overhead powerline network and stepped down at five surface substations, one serving each of the five underground portal areas (Portal 1 to Portal 5). Medium voltage (5.5 kV) will then be distributed underground to skid-mounted substations that supply 1,000 V power to the underground loads.

16.9.1Surface Electrical Infrastructure

Five surface substations will transform the incoming 22 kV supply to 5.5 kV. Three substations will be rated at 2 MVA (Portals 1, 2 and 5) and two at 1.5 MVA (Portals 3 and 4), for a total installed surface transformer capacity of 9 MVA. The substations will be fed from the surface 22 kV overhead powerline network, and each will be capable of supplying the underground ventilation fans and all underground levels within its portal area. Table 16-25 summarizes the surface substations.

All surface substations will be modular, housed in 40-ft high-cube containers, so that they can be relocated as required by the development sequence. All underground communication cabinets and devices will be installed in the same surface containers.

Table 16-25    Surface Substations by Portal

Portal			Transformation			Capacity (MVA)			Levels Served
Portal 1			22 kV – 5.5 kV			2			27
Portal 2			22 kV – 5.5 kV			2			21
Portal 3			22 kV – 5.5 kV			1.5			18
Portal 4			22 kV – 5.5 kV			1.5			25
Portal 5			22 kV – 5.5 kV			2			30
Total						9			121

One 5.5 kV feeder, using riser TECK cable, will be installed through each portal to supply the underground substations on the first levels. Once the ventilation raise is completed, a second 5.5 kV feeder will be installed through the raise to supply the lower levels.

16.9.2Emergency Power

A common modular diesel generator station will be installed near the surface substations to supply the underground emergency loads, including compressors, pumps and ventilation. The generator station grounding will be connected to the isolated grounding grid to maintain a constant fault level. Fuel storage capacity will be sufficient to supply the emergency loads for 24 hours.

16.9.3Underground Electrical Distribution

Junction boxes will be installed on each level to feed the underground main substations. The underground main substations will be installed at each main level and will provide 1,000 V power to underground loads such as pumps, ventilation and garages. The key design features of the underground distribution system are as follows:

•One 1 MVA, 5.5 kV – 1 kV underground substation is planned for every three levels;

•Substations are expected to be skid-mounted to allow ease of installation and relocation;

•All substation equipment will be located inside an electrical room excavation along the main ramp;

•As underground mine development progresses, some substations will be relocated from completed areas to newly developed zones.

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16.9.4Communications

The mine will be provided with a fibre optic backbone and a private LTE (Long Term Evolution) wireless network to support all communications required to operate the mine and to transmit data within the mine and to surface. The communication infrastructure will support voice communications, PLC monitoring and control, video, operational data, and control and monitoring of the electrical network.

The fibre optic backbone will be installed from the surface control room through each portal and down the ramps and ventilation raises, with connections to the electrical substations and to each active level. It will carry all wired Ethernet/IP traffic and serve as the transport network for the LTE system. A surface LTE core and backbone will link the underground network to the site operations centre. The fibre backbone will be installed in a ring configuration where practical to provide redundancy and improved reliability.

LTE radio units will be installed along the ramps, level accesses and production drifts as development advances, and connected to the fibre backbone. The LTE network will provide high-bandwidth, low-latency wireless coverage for voice, data and video across the underground workings. The network will also be capable of supporting additional applications such as personnel and equipment tracking, real-time equipment telemetry, and tele-remote operation of mobile equipment. These applications are not included in the base case but may be implemented during operations, providing flexibility to improve productivity and safety without major changes to the communication infrastructure.

A dedicated blast line communication system will also be installed to support blasting operations.

16.9.5Electrical Infrastructure Unit Costs

The unit costs provided for the main electrical and communication infrastructure are summarized in Table 16-26. Quantities and total capital costs are addressed in the capital cost estimate.

Table 16-26    Electrical and Communication Infrastructure Unit Costs

Item			Description			Unit Cost (US$)
Surface substation			2 MVA, 22 kV – 5.5 kV			600,000
Surface substation			1.5 MVA, 22 kV – 5.5 kV			500,000
Underground substation			1 MVA, 4.16 kV – 1 kV, skid-mounted			250,000
Communication system			LTE			500,000
Blast line communication			Blast line communication system			100,000

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16.10Ventilation

This section of the report outlines the ventilation design and strategy for the Boumadine Project, based on the proposed mine layout, production schedule, and diesel equipment fleet. The design aims to ensure safe and efficient airflow distribution across all zones of the mine, while optimizing energy consumption and fan performance. Air flow rate is based on a diesel dilution rate of 0.06 m³/s/kW as per industry standard.

16.10.1Design Criteria

Other design criteria for ventilation are defined as follows:

•Maximum air velocity in ramps: 7 m/s

•Main fan pressure limit: 5000 Pa to avoid high pressures in return air raise system for safety and bulkhead design purposes to access the egress.

•Fan efficiency of main fans: 75%

•Fan efficiency of auxiliary fans: 80%

•Main fan electrical utilization rate: 80%.

•Motor efficiency: 95%

•Fan installation loss pressure: 625 Pa

•Heating or cooling of mine air is not required.

•Any open stopes are sealed with a ventilation curtain

•Optimize blast clearing times

•Leakage factor: 50 mm²/m² for auxiliary duct network

•Hydraulic resistance of closed regulator and personnel doors bulkheads: 15 N s2/m8

•When airflow is divided between two raises on a given zone, a 70/30% split is assumed for worst-case pressure balance.

•Electricity cost: $US 90/MWh

•Main fans operate 365 days/year, 24 hours/day

•Auxiliary fans operate 365 days/year, 24 hours/day with On/Off starters only

•Vertical mining cost of $US 1537/m plus 218/m3

16.10.2Ventilation Demand and Distribution

Ventilation demand is based on the number of equipment and utilization factor as shown in Table 16-27, specified before the addition of 15% leakage.

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Table 16-27    Diesel equipment air demand

Machine			Engine (kW)			m3/s/kW			Utilisation (%)			Airflow per unit, (m3/s)			No. of units			Total airflow (m3/s)
Drill DL311			74.6			0.06			25			1.1			3			3.4
Jumbo			74.6			0.06			25			1.1			9			10.1
Scissor Paus			111.9			0.06			50			3.4			9			30.2
LHD 10t 1030			186.5			0.06			100			11.2			10			111.9
Bolter Sandvik DS421			74.6			0.06			25			1.1			9			10.1
Emulsion PAUS			111.9			0.06			25			1.7			3			5.0
Truck 33t MT436B			298.4			0.06			100			17.9			6			107.4
Subtotal 1 - Allocated to resources																		278
			111.9			0.06			25			1.7			1			1.7
Transmixer			111.9			0.06			50			3.4			3			10.1
Boom truck Paus			111.9			0.06			50			3.4			1			3.4
Grader GR135 XCMG			111.9			0.06			50			3.4			2			6.7
Fuel Lube Paus			111.9			0.06			50			3.4			1			3.4
Water Truck Paus			93.25			0.06			50			2.8			15			42.0
Subtotal 2 - Distributed Proportionally to Resources																		67

Airflow is allocated to resources defined in Table 16-28. A production schedule from 2029 to 2044 specifies the number of resources in each zone. The remaining fleet air demand is distributed proportionally to the resource allocation, with an additional amount of 15% for leakage. The resulting peak airflow demand in each zone is shown in Table 16-29. The airflow demand for the life-of-mine by zone is shown in Figure 16-33.

Table 16-28    Task resource definitions

Resource category			Resources assigned			Airflow, m3/s
Dev Horizontal Capex			1 Jumbo 1 Scissor lift ½ LHD 1 Cable bolter ¼ LHD ¼ emulsion truck			8.8
Dev Horizontal Opex			1 Jumbo 1 Scissor lift ½ LHD 1 Cable bolter ¼ LHD ¼ emulsion truck			8.8
Dev Vert			½ LHD			5.6
Drilling			Production drill and compressors			1.5
Mucking			1 LHD			11.2
Backfilling			1 LHD			11.2
Haulage			1 Truck and ¼ LHD			20.7

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Table 16-29    Peak airflow by zone, including 15% leakage

Zone			Peak airflow (m3/s)			Start of production
Portal 1			255			2031
Portal 2			95			2036
Portal 3			133			2036
Portal 4			226			2034
Portal 5			285			2038

Figure 16-33    Life of mine airflow by zone Portal 1 to Portal 5

16.10.3Main Ventilation system Design

The main ventilation system design for all zones consists of the following:

•Fresh air will enter the mine via portal and dedicated intake raises

•Fresh air will descend via ramp to the levels

•Fresh air will be supplied to the working faces by auxiliary ducting and fans

•Exhaust air will enter dedicated return air raises by drop-board regulators

•Air will be exhausted from the mine by primary fans installed on surface.

This design was chosen to optimize blast clearing times. The drop board regulators are adjusted manually to supply enough air on each of the levels. When a level becomes inactive, the regulator is completely sealed.

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Figure 16-34    Zones and locations of exhaust fans

The extent of the underground mining zones is pictured in Figure 16-34.

Zone South 1, pictured in Figure 16-35, features three fan installations which will be sized to exhaust, respectively, 50%, 75%, and 50% of the maximum air flow for the zone, that is 255 m3/s reached in year 2034.

Figure 16-35    Longitudinal view of South 1 zone showing three exhaust raises and three intakes

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Figure 16-36    Longitudinal view of South 2 zone showing exhaust raise and intake portal

Zone South 2 will require a maximum of 95 m3/s, that will be provided by a single exhaust fan, shown in Figure 16-36.

Figure 16-37    Longitudinal view of South 3 zone showing three exhaust raises and three intakes

Zone South 3, pictured in Figure 16-37, features three fan installations which will be sized to exhaust, respectively, 100%, 35%, and 35% of the maximum air flow for the zone, that is 133 m3/s reached in year 2036.

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Figure 16-38    Longitudinal view of Center (Portal 4) zone showing two exhaust raises and two intakes

Zone Center (Portal 4), pictured in Figure 16-38, features two fan installations which will be sized to exhaust, respectively, 50%, and 100% of the maximum air flow for the zone, that is 226 m3/s reached in year 2034.

Figure 16-39    Longitudinal view of North (Portal 5) zone showing four exhaust raises and four intakes

Zone North (Portal 5), pictured in Figure 16-39, features four fan installations which will be sized to exhaust, respectively, 70%, 70%, 40%, and 40% of the maximum air flow for the zone, that is 285 m3/s reached in year 2038.

Thirteen main fan installation locations are listed in Table 16-30. Some of the smaller fans may be specified by vendors as vertically mounted for reduced cost. Raises exhausting more than 150 m3/s of air will more likely be quoted for twin parallel horizontal fans.

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Table 16-30    Main fan operating points and installation configuration

Portal			Fan			Maximum flow as % of zone			Max. airflow (m3/s)			Fan pressure (Pa)			Fan type			Combined power (kW)			Fans
Portal 1			South 1.1			50			102			3000			Surface horizontal			500			1
Portal 1			South 1.2			75			179			3600			Surface horizontal parallel arrangement			900			2
Portal 1			South 2.1			50			102			2300			Surface horizontal or vertical			400			1
Portal 2			South 2.2			100			95			2200			Surface horizontal or vertical			300			1
Portal 3			South 3			100			133			2700			Surface horizontal or vertical			500			1
Portal 3			South 4.1			35			80			2300			Surface horizontal or vertical			250			1
Portal 3			South 4.2			35			80			2000			Surface horizontal or vertical			200			1
Portal 4			Center			100			226			4300			Surface horizontal parallel arrangement			1300			2
Portal 4			South 5			50			113			3000			Surface horizontal or vertical			450			1
Portal 5			North 1			70			199			3200			Surface horizontal parallel arrangement			850			2
Portal 5			North 2			40			114			2400			Surface horizontal or vertical			400			1
Portal 5			North 3.1			40			114			2900			Surface horizontal or vertical			450			1
Portal 5			North 3.2			70			199			4500			Surface horizontal parallel arrangement			1200			2

Vertical raise sizing is based on prices for 3-m ($2,525/m) and 4-m ($3,000/m) diameter raise resulting in an variable cost of $1945/m and 83/m3. Raises to surface are raisebore specified up to 4.0 m diameter maximum to reduce ground control risk during excavation. All underground ventilation raises are drop raises at a maximum size of 4.5 × 4.5 m and minimum of 3.0 m x 3.0 m. The RAR system is also used as an egress using laddertube type. The laddertube is maintained open to minimize the impact on the hydraulic resistance of the RAR system. Double personnel doors are therefore installed in each of the bulkheads for egress purposes. A summary of vertical airways according to dimension and type of excavation is provided in Table 16-31.

Table 16-31    Summary of raises by dimensions and length

Raise type			Total length, m			Total number of raises
Drop Raise 3.0x3.0			2261			113
Drop Raise 3.5x3.5			461			23
Drop Raise 4.0x4.0			1033			52
Drop Raise 4.5x4.5			560			28
Raisebore 3.0m			1568			6
Raisebore 3.65m			1435			10
Raisebore 4.0m			1496			8

Main fan power costs are estimated in Table 16-32 assuming an average power utilization of 80% which is equivalent to an average speed reduction of 8%. Speed reductions are possible with the use of Variable Frequency Drives (VFDs).

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Table 16-32    Energy Consumption of Main Fans

Zone			Fan power (kW)			Utilisation			2031			2032			2033			2034			2035			2036			2037			2038			2039			2040			2041			2042			2043
Portal 1			1800			80%			1.1$			1.1$			1.1$			1.1$			1.1$			1.1$
Portal 2			300			80%																					0.2$			0.2$			0.2$			0.2$
Portal 3			950			80%																		0.6$			0.6$			0.6$			0.6$
Portal 4			1750			80%												1.1$			1.1$			1.1$			1.1$			1.1$			1.1$			1.1$			1.1$			1.1$
Portal 5			2900			80%																								1.8$			1.8$			1.8$			1.8$			1.8$			1.8$
Total, $M									1.1$			1.1$			1.1$			2.2$			2.2$			2.8$			1.9$			3.7$			3.7$			3.1$			2.9$			2.9$			1.8$
Annual MWh									12,614			12,614			12,614			24,878			24,878			31,536			21,024			41,347			41,347			34,690			32,587			32,587			20,323

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16.10.4Auxiliary Ventilation system Design

Auxiliary ventilation for production headings and ramp advance was sized using a duct leakage allowance of 50 mm²/m². Fan and motor efficiencies of 65% and 95%, respectively, were applied when establishing the required duty points. The assessment also accounts for the duct lengths associated with the planned advance rates and the geometry of the production levels.

16.10.4.1Ramp Development

Ramp headings will be ventilated through two 1.0-m-diameter flexible duct lines, each designed for runs of up to 800 m. Fresh air will be supplied from the main ventilation network, with return-air-raise (RAR) connections established at five-level intervals. Each duct system is equipped with a 200-kW auxiliary fan designed to deliver 37 m3/s while supporting the concurrent operation of a truck, LHD, bolter and pickup. When a new RAR connection becomes available, the fans and ducting will be advanced to the lower connection. The mine plan assumes that two ramp-development ventilation systems may operate simultaneously. A typical ramp-development arrangement is presented in Figure 16-40.

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Figure 16-40    Auxiliary Ventilation on Ramp Development Zone Centre

16.10.4.2Production Levels

Ventilation on each active production level will be provided by a 150-kW main auxiliary fan, supplemented by a 15-kW fan serving secondary headings. Fresh-air quantities will be controlled at the level regulator to satisfy fan intake requirements, prevent recirculation and maintain adequate airflow to the other active levels. The design basis allocates 21 m³/s to the primary heading, where an LHD and truck may operate, and 10 m³/s. The lower airflow allowance applies to a secondary heading served by an LHD. Both circuits use 1.0-m-diameter flexible ducting, with representative lengths of 300 m for primary headings and 200 m for secondary headings. At the mine-wide peak, eleven levels are assumed to have fan installations: four operating, four available on standby and three supporting development. The duct-leakage allowance of 50 mm²/m² is included when determining fan pressure and the airflow delivered to the working face. A typical production-level ventilation arrangement is presented in Figure 16-41.

Figure 16-41    Auxiliary Ventilation on Production Levels Zone Centre

The combined annual energy cost of the production-level and ramp-development auxiliary fans is estimated at US$1.9 million per year, based on an electricity rate of US$90/MWh. The corresponding fan duties and energy calculations are summarized in Table 16-33.

Table 16-33    Energy Consumption of Auxiliary and Ramp Development Fans

			Airflow (m3/s)			Pressure (Pa)			Consumed power per fan (kW)			Average number of fans operating			Annual energy consumption (MWh)			Annual energy cost ($US)
Production level main auxiliary fan			21			4117			140			11			13,491			1,214,234
Production level secondary fan			10			656			11			11			1,024			92,131
Ramp development fan			18.5			6364			191			4			6,681			601,273
												Total			21,196			1,907,638

The combined annual energy cost for auxiliary and main ventilation systems is estimated on average at $US 4.2 M/yr.

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17.0RECOVERY METHODS

17.1Process Design

The process plant design for the Project is based on a conventional flotation flowsheet. Lead, zinc and pyrite concentrates will be recovered and sold to market.

The plant design criteria were selected to minimize operating costs and maximize the use of proven technology. The key criteria for equipment selection are suitability for duty, safety, reliability, and ease of maintenance.

17.1.1Selected Process

The process plant design comprises the following unit operations:

•Single stage primary crushing with a jaw crusher to produce an 80% Passing (P80) crushed product size of 136 mm.

•Crushed rock stockpile with a live capacity of approximately 8,115 tonnes to provide 24 hours of live storage. During extended periods of primary crusher equipment maintenance, additional crushed material inventory can be generated in the weeks leading up to the planned shutdown by dozing crushed material from this stockpile to the area adjacent to the stockpile. This material can then be reclaimed during the shutdown by front-end loader to feed the grinding circuit.

•SAB type grinding circuit consisting of a SAG mill and a ball mill to produce a ground product with a P80 of 58 microns (µm), with hydrocyclones for particle size classification.

•Lead rougher flotation, classification, regrinding and cleaner flotation to produce a lead concentrate of 29.6% grade. The lead concentrate will be thickened, filtered, and bagged prior to shipment.

•Zinc rougher flotation, classification, regrinding and cleaner flotation to produce a zinc concentrate of 57.4% grade. The zinc concentrate will be thickened and filtered for stockpiling prior to shipment.

•Pyrite rougher flotation to produce a pyrite concentrate with 4.2 Au g/t and 81.0 Ag g/t. The pyrite concentrate will be thickened and filtered for stockpiling prior to shipment.

•Tailings thickening and storage in a tailings management facility.

The unit operations incorporated in the selected process flowsheet are shown in Figures 17-1 and 17-2.

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Figure 17-1    Simplified Process Flowsheet

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Figure 17-2    Simplified Process Flowsheet (Continued)

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17.1.2Key Process Design Criteria

The key design criteria for the process plant are outlined in Table 17-1.

Table 17-1    Key Process Design Criteria

Criteria Description			Units			Nominal			Source
Plant Throughput			Mtpa			2.92			AGS
Life of Mine			y			11			AGS
Head Grade - Design			g/t Au			3.15			Testwork
			g/t Ag			105			Testwork
			% Pb			0.92			Testwork
			% Zn			1.87			Testwork
Flotation Recovery: •Pb to Pb Concentrate •Zn to Zn Concentrate •Au to Py Concentrate •Ag to Py Concentrate			% % % %			82.0 74.7 71.4 41.4			Testwork Testwork Testwork Testwork
Concentrate Grades: •Pb grade in Pb Concentrate •Zn grade in Zn Concentrate •Au grade in Py Concentrate •Ag grade in Py Concentrate			% % g/t g/t			29.6 57.4 4.2 81.0			Testwork Testwork Testwork Testwork
Reagent Addition1: •Zinc Sulphate (ZnSO4·7H2O) •Sodium-Diisobutyl Dithiophosphinate (Aerophine 3418A) •Methyl Isobutyl Carbinol (MIBC) •Copper Sulphate (CuSO4·5H2O) •Sodium Isopropyl Xanthate (SIPX) •Potassium Amyl Xanthate (PAX) •Sulphuric Acid (H2SO4) •Sodium Cyanide (NaCN) •Lime (@91% CaO)			g/t g/t g/t g/t g/t g/t g/t g/t g/t			360 22.5 106.3 225 32.5 300 1,000 120 2,350			Testwork Testwork Testwork Testwork Testwork Testwork Testwork Testwork Testwork
Crushing Plant Availability			%			75			Agreed
Milling and Flotation Circuit Availability			%			91.3			Agreed
Filtration Circuit Availability			%			80			Agreed
Crushing Work Index (CWi)			kWh/t			10.2			Assumption
Bond Ball Mill Work Index (BWi)			kWh/t			13.1			Testwork
SMC Axb			0			46.2			Assumption
Bond Abrasion Index (Ai)			0			0.402			Assumption
Grind Size, P80			µm			58			Testwork
Pb Flotation Conc. Mass Pull – Rougher / Cleaner			%			9.7 / 2.6			Testwork
Pb Regrind Size, P80			µm			16			Testwork
Pb Concentrate Thickener Solids Loading			t/h/m2			0.2			Assumption
Zn Flotation Conc. Mass Pull – Rougher / Cleaner			%			4.4 / 2.4			Testwork
Zn Regrind Size, P80			µm			15			Testwork
Zn Concentrate Thickener Solids Loading			t/h/m2			0.2			Assumption
Py Rougher Flotation Conc. Mass Pull			%			53.5			Testwork
Py Concentrate Thickener Solids Loading			t/h/m2			0.5			Assumption
Flotation Tailings Thickener Solids Loading			t/h/m2			0.5			Assumption
Tailings Disposal			-			Pumped to TSF (by others)			AGS

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17.2Process Plant Description

ROM material will be delivered to a local feed bin by a mine truck to feed the crushing plant. The crushed product will be stockpiled prior to feeding the grinding circuit, consisting of a SAG mill and a ball mill with hydrocyclones for particle size classification. The ground product will feed the lead rougher flotation conditioning tank, which will feed the lead rougher flotation cells. The lead rougher flotation concentrate will proceed to a classification and regrind circuit, while the tails will feed the zinc rougher flotation conditioning tanks. Following regrinding, the ground lead rougher concentrate will report to a lead cleaner flotation circuit, consisting of three cleaning stages and one cleaner-scavenger stage. The lead cleaner tailings will report to zinc rougher flotation, while the final lead cleaner concentrate will be thickened, filtered and bagged prior to shipment. The filtrate from the lead filter will circulate back to the lead concentrate thickener and the lead thickener overflow will report to the lead water circuit and any excess will report back to the process water pond.

The lead rougher tails and the lead first cleaner-scavenger tails will feed the zinc rougher flotation conditioning tank, which will feed the zinc rougher flotation cells. The zinc rougher flotation concentrate will proceed to a classification and regrind circuit, while the tails will feed the pyrite rougher flotation conditioning tanks. Following regrinding, the zinc rougher concentrate will report to a zinc cleaner flotation circuit, consisting of three cleaning stages and one cleaner-scavenger stage. The zinc cleaner-scavenger tailings will report to pyrite rougher flotation, while the third zinc cleaner concentrate will be thickened and filtered prior to stockpiling for truck loading and shipment. The filtrate from the zinc filter will circulate back to the zinc concentrate thickener and the zinc thickener overflow will report to the zinc water circuit and any excess will report back to the process water pond.

The zinc rougher tails and the zinc first cleaner-scavenger tails will feed the pyrite rougher flotation conditioning tank, which will feed the pyrite rougher flotation cells. The pyrite rougher flotation tailings will be thickened in a tailings thickener before getting pumped to the flotation tailings management facility. The flotation tailings thickener overflow will report to the process water pond. The pyrite concentrate will be thickened, filtered and stockpiled by a radial stacker from where it is loaded on trucks and shipped. The filtrate from the pyrite concentrate filter will report to the pyrite concentrate thickener, and the pyrite concentrate thickener overflow will report to the process water pond.

17.2.1Crushing Circuit

The crushing circuit will be fed through the ROM bin by direct tipping of the mine truck or reclaimed by a front-end loader. The ROM bin will be fitted with a static grizzly with an aperture size of 800x800 mm to prevent oversized rocks from entering the crusher. Oversized rocks will be crushed by a fixed rock breaker. The ROM bin will discharge onto an apron feeder that will feed a vibrating grizzly with an aperture size of 90 mm. The grizzly undersize will bypass the primary crusher to reduce wear on the crusher, while the grizzly oversize will directly feed the primary crusher. The primary crusher will be a 160 kW jaw crusher, and will crush the grizzly oversize at a rate of 444 dry t/h from an F80 of 501 mm to a P80 of 156 mm. The crushed rock will discharge onto the primary crusher discharge conveyor together with the grizzly undersize to feed the 24h live capacity crushed rock stockpile.

17.2.2Grinding Circuit

Crushed material will be reclaimed from the crushed rock stockpile by two apron feeders, which may be operated together or with one on standby. The apron feeders will discharge on to the SAG mill feed conveyor to feed the primary mill. Flotation reagents zinc sulphate, sodium cyanide and Aero 3418A will be added to the primary mill with process water to pre-condition the material and to achieve a target moisture level of 75 %w/w in the mill.

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The primary mill will be a SAG mill with a diameter of 7.92 m and effective grinding length of 4.39 m. It will have a high-speed variable speed drive and installed motor power of 6.1 MW. It will be charged with forged steel balls with a maximum size of 125 mm, equipped with a grate discharge and a trommel screen. The trommel screen oversize will discharge into a scats bunker or onto one of two pebble transfer conveyors for recirculation back to the SAG mill via the SAG mill feed conveyor. The trommel screen undersize will discharge to the cyclone feed pump box. The trommel screen will be equipped with a spray water mechanism to improve screening efficiency.

The secondary mill will be a ball mill with a diameter of 6.10 m and an effective grinding length of 9.32 m. Process water will be fed to the mill to achieve a target moisture content of 75%. The ball mill will have a fixed speed drive with a wound round induction motor and liquid resistance starter. The installed motor power will be 6.1 MW. It will be charged with high-chromium steel balls with a maximum size of 50 mm, discharging by overflow and equipped with a trommel screen. The trommel screen oversize will discharge to a scats bunker, and the trommel screen undersize will discharge to the cyclone feed pump box. The trommel screen will be equipped with a spray water mechanism to improve screening efficiency.

The combined SAG and Ball Mill discharge will be pumped to the cyclones for classification from the cyclone feed pump box, targeting an overflow P80 of 58 µm. There will be a total of twelve installed cyclones with six duty and six standby. The cyclones will be operated at a pressure of 85 kPa(g) and are expected to have a nominal recirculation rate of 382%. The cyclone overflow will report to the lead rougher flotation circuit, and the underflow will feed the secondary mill.

17.2.3Lead Rougher Flotation Circuit

The cyclones’ overflow will gravitate to the lead rougher conditioning tank. Zinc sulphate, sodium cyanide and Aero 3418A will be fed to the agitated conditioning tank where the rougher feed will be conditioned for five minutes. The conditioned slurry will then feed the lead rougher flotation cells. The circuit will consist of a single train of six forced-air mechanical cells. Zinc sulphate, sodium cyanide, Aero 3418A, and MIBC will be added to the lead rougher cells to maximize lead recovery. The lead rougher flotation concentrate will report to the lead regrind and cleaner flotation circuits, while the tails will feed the zinc rougher flotation circuit.

17.2.4Lead Regrind Circuit

The lead rougher concentrate and the lead first cleaner-scavenger concentrate will report to the lead regrind cyclone pump box and will be pumped to the lead regrind cyclone cluster for classification and to increase the percent solids of the slurry feeding the regrind mill. The lead cyclone cluster will consist of sixteen cyclones in total, with 11 duty and 5 standby. The cyclones will be operated at a pressure of 64 kPa(g) and will target an overflow particle d50 of 21 µm. The lead regrind cyclones’ overflow will report directly to the lead cleaner flotation surge tank, while the underflow will gravitate to the lead regrind mill feed pump box before getting pumped to the lead regrind stirred media mill (SMM). The SMM will reduce the particle size from an F80 of 58 µm to a P80 of 16 µm. It will be charged with zirconia toughened alumina ceramic beads and will have an installed motor power of 1.6 MW. The SMM discharge will combine with the cyclone overflow in the lead cleaner flotation surge tank.

17.2.5Lead Cleaner Flotation Circuit

Zinc sulphate and sodium cyanide will be added to the lead cleaner flotation surge tank for conditioning ahead of flotation. The flotation feed will be conditioned for five minutes prior to feeding the first lead cleaner flotation circuit. The lead cleaner flotation circuit will consist of three cleaning stages and one cleaner-scavenger stage. The first cleaner stage will contain one train of four forced-air mechanical cells and all other stages will contain one train of three forced-air mechanical cells. MIBC will be added to each stage and Aero 3418A will be added to the second and third cleaner flotation circuits to promote lead recovery.

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The concentrate and tails will travel through the circuit counter-currently. The lead first cleaner flotation concentrate will feed the lead second cleaner flotation cells, the lead second cleaner flotation concentrate will feed the lead third cleaner flotation cells, and the lead third cleaner flotation concentrate will feed the lead concentrate thickener. The lead third cleaner flotation tails will report to the lead second cleaner flotation cells, the lead second cleaner flotation tails will feed the lead first cleaner flotation cells, and the lead first cleaner flotation cell tails will feed the lead first cleaner-scavenger flotation cells. The lead first cleaner-scavenger flotation concentrate will be recirculated back to the lead regrind cyclone pump box, while the tails will report to the zinc rougher flotation conditioning tank with the lead rougher flotation tails.

17.2.6Lead Concentrate Thickening, Filtration, and Bagging Circuit

The lead third cleaner flotation concentrate will be pumped to the lead concentrate thickener, where flocculant will be added to the feed well to promote particle settling. The thickener will be high rate with a diameter of 11 m. The slurry will be thickened to 55% (w/w) solids in the underflow. The thickener overflow will report to the lead process water tank, while the thickener underflow will be pumped to the lead concentrate filter feed tank. The filter feed tank will be sized to hold 24 hours of feed slurry. The slurry will be pumped to a pressure filter to dewater the concentrate cake to 10% (w/w) moisture. The thickener and filter design criteria will be further defined upon the completion of dewatering testwork. The filtrate from the lead concentrate filter will be discharged to a filtrate tank and pumped back to the lead concentrate thickener, while the lead concentrate filter cake will be discharged via a belt feeder, bucket elevator and shuttle conveyor into lead concentrate hoppers. The hoppers will feed automatic bagging systems, with the bags transported and stacked by forklift prior to loading on highway trucks for shipment offsite.

17.2.7Zinc Rougher Flotation Circuit

The first zinc rougher conditioning tank will receive the lead rougher flotation tails and the lead first cleaner scavenger tails. Lime will be added to the first conditioning tank to achieve a target pH of 11.5, and the second conditioning tank will be dosed with additional lime and copper sulphate. The zinc rougher flotation feed will be conditioned in the agitated conditioning tanks for a total of 10 minutes – five minutes in each tank. The conditioned slurry from the second conditioning tank will feed the zinc rougher flotation cells. The circuit will consist of a single train of six forced-air mechanical cells. SIPX, lime and MIBC will be added to the zinc rougher cells to maximize zinc recovery. The zinc rougher flotation concentrate will report to the zinc regrind and cleaner flotation circuits, while the tails will feed the pyrite rougher flotation circuit.

17.2.8Zinc Regrind Circuit

The zinc rougher concentrate and the zinc first cleaner-scavenger concentrate will report to the zinc regrind cyclone pump box and will be pumped to the zinc regrind cyclone cluster for classification and to increase the percent solids of the feed entering the regrind mill. The zinc cyclone cluster will consist of eight cyclones in total, with 5 duty and 3 standby. The cyclones will be operated at a pressure of 81 kPa(g) and will target an overflow particle d50 of 20 µm. The zinc regrind cyclones overflow will report directly to the zinc cleaner flotation surge tank, while the underflow will gravitate to the zinc regrind mill feed pump box before being pumped to the zinc regrind stirred media mill (SMM). The SMM will reduce the particle size from an F80 of 58 microns to a P80 of 15 µm. It will be charged with zirconia toughened alumina ceramic beads and will have an installed motor power of 700 kW. The SMM discharge will combine with the cyclone overflow in the zinc cleaner flotation surge tank.

17.2.9Zinc Cleaner Flotation Circuit

SIPX and lime will be added to the zinc cleaner flotation surge tank for conditioning ahead of flotation. The flotation feed will be conditioned for five minutes prior to feeding the first zinc cleaner flotation circuit. The

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zinc cleaner flotation circuit will consist of three cleaning stages and one cleaner-scavenger stage. Each stage will contain one train of three forced-air mechanical cells. MIBC will be added to each stage, SIPX will be added to the first cleaner-scavenger and the third cleaner, and lime will be added to the first through third cleaners to maximize zinc recovery.

The concentrate and tails will travel through the circuit counter-currently. The zinc first cleaner flotation concentrate will feed the zinc second cleaner flotation cells, the zinc second cleaner flotation concentrate will feed the zinc third cleaner flotation cells, and the zinc third cleaner flotation concentrate will feed the zinc concentrate thickener. The zinc third cleaner flotation tails will report to the zinc second cleaner flotation cells, the zinc second cleaner flotation tails will be recycled to the zinc first cleaner flotation cells, and the zinc first cleaner flotation cell tails will report to the zinc first cleaner-scavenger flotation cells. The zinc first cleaner-scavenger flotation concentrate will be recirculated back to the zinc regrind cyclone pump box, while the tails will report to the pyrite rougher flotation conditioning tank with the zinc rougher flotation tails.

17.2.10Zinc Concentrate Thickening and Filtration Circuit

The zinc third cleaner flotation concentrate will be pumped to the zinc concentrate thickener, where flocculant will be added to the feed well to promote particle settling. The slurry will be thickened in a 11 m diameter, high-rate thickener to 55% (w/w) solids in the underflow. The thickener overflow will report to the zinc process water tank, while the thickener underflow will be pumped to the zinc concentrate filter feed tank. The filter feed tank will be sized to hold 24 hours of feed slurry. The slurry will be pumped to a pressure filter to dewater the concentrate cake to 10% (w/w) moisture. The thickener and filter design criteria will be further defined upon the completion of dewatering testwork. The zinc concentrate filter filtrate will be discharged to filtrate tank and pumped back to the zinc concentrate thickener, while the zinc concentrate filter cake will be discharged to a stockpile in a covered shed prior to loadout. The zinc concentrate will be loaded into highway haulage trucks by a front-end loader for shipment offsite.

17.2.11Pyrite Rougher Flotation Circuit

The pyrite rougher flotation conditioning tank will receive the tails from the zinc rougher flotation and the zinc first cleaner-scavenger flotation. Sulphuric acid will be fed to the agitated conditioning tank to lower the pH between 7-8.5. The pyrite rougher feed will be conditioned for five minutes. The conditioned slurry will feed the pyrite rougher flotation cells. The circuit will consist of a single train of six forced-air mechanical cells. PAX and MIBC will be added to the pyrite rougher cells to maximize pyrite recovery. The pyrite rougher flotation concentrate will report to pyrite concentrate thickening and filtration, while the tails will report to the flotation tailings thickener.

17.2.12Pyrite Concentrate Thickening and Filtration Circuit

The pyrite rougher flotation concentrate will be pumped to the pyrite concentrate thickener, where flocculant will be added to the feed well to promote particle settling. The slurry will be thickened to 65% (w/w) solids in the underflow in a 32 m diameter high-rate thickener. The thickener design criteria will be further defined upon the completion of dewatering testwork. The thickener overflow will report to the process water pond, while the thickener underflow will be pumped to the pyrite concentrate filter feed tank. The filter feed tank will be sized to hold 24 hours of feed slurry. From the filter feed tank, the pyrite concentrate slurry is pumped to one of two parallel pressure filters to dewater the concentrate cake to 15% (w/w) moisture for stockpiling.

The filtrate from the pyrite concentrate filter will recirculate back to the pyrite concentrate thickener. The filter design criteria will be further defined upon the completion of dewatering testwork. The stockpile will be a radially stacked dead stockpile with a live capacity of 7,250 m3, holding approximately 24 hours of filtered pyrite concentrate. The pyrite concentrate will be loaded into highway haulage trucks by a front-end loader for shipment offsite.

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17.2.13Flotation Tailings Disposal

The pyrite rougher flotation tails will be pumped to the flotation tails thickener, where flocculant will be added to the feed well to promote particle settling. The thickener will be a high-rate type with a diameter of 22 m. The slurry will be thickened to 40% (w/w) solids in the underflow. The thickener overflow will report to the process water pond, while the thickener underflow will be pumped to the flotation tailings management facility. It is estimated that 129 m3/h of water will be decanted from the flotation tailings management facility back to the process water pond. The thickener design criteria will be further defined upon the completion of dewatering testwork.

17.2.14Reagents

Plant operations will require a reagent inventory of flocculant, zinc sulphate, Aero 3418A, MIBC, copper sulphate, SIPX, PAX, sulphuric acid, sodium cyanide, lime, and anti-scalant. The site will maintain six weeks’ worth of supply of each reagent onsite in the reagent shed. All reagents will be dosed to their respective injection sites by metered dosing pumps (one per dosing site), except for lime and flocculant. Lime will be dosed by a centrifugal pump and flocculant will be dosed by helical screw pumps, with one per dosing application.

17.2.14.1Flocculant

Flocculant will be used in each of the thickeners to promote particle settling. It will be delivered to site as a solid powder in 750 kg bags. The bags will be split by a bag breaker and mixed for five hours with filtered process water. The mixing concentration will be 0.25% (w/v). The flocculant will be diluted in-line with process water to a concentration of 0.025% (w/v). The type and quantity of flocculant used will be further defined upon the completion of thickening testwork.

17.2.14.2Zinc Sulphate

Zinc sulphate is an activator in lead flotation. It will be fed to the SAG mill, lead rougher flotation and lead cleaner flotation feed. It will be delivered to site as a heptahydrate crystal in 1,000 kg bags at 98% purity. It will be diluted to 25% (w/v) concentration for dosage. Three bags will be required for the mixing of one batch, which is enough to supply 22 hours of operation.

17.2.14.3Aero 3418A

Aero 3418A will be used as a collector and promoter in lead flotation. It will be fed to the SAG mill, lead rougher flotation, the lead first cleaner and the lead first cleaner scavenger. It will be delivered to site and stored in 1 m3 IBC totes.

17.2.14.4MIBC

MIBC will be delivered as a liquid in 1 m3 IBC totes for usage as a frothing agent in all flotation circuits and stages.

17.2.14.5Copper Sulphate

Copper sulphate will be used as an activator in zinc rougher flotation. It will be fed to the second zinc conditioning tank. The copper sulphate will be delivered as solid pentahydrate crystals at 98% purity in 1,200 kg bulk bags. The copper sulphate will be diluted to 20% (w/v) for dosage. Two tonnes of crystals will be required per mix, which is enough to supply operations for 24 hours.

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17.2.14.6Sodium Isopropyl Xanthate (SIPX)

SIPX will be used as a collector in zinc rougher and cleaner flotation. It will be fed to the zinc rougher, zinc regrind feed, zinc first and second cleaners, and the zinc first cleaner scavenger. SIPX will be delivered to site in 1,000 kg bulk bags as pellets with a purity of at least 90%. One tonne of material will be used per mix, through which the SIPX will be diluted to 10% (w/v). One mix will last for approximately three days of operation.

17.2.14.7Potassium Amyl Xanthate (PAX)

PAX will be used as a collector in the pyrite rougher flotation, fed directly to the rougher cells. PAX will be delivered to site in pellet format packed in 1,000 kg bulk bags, with a purity of at least 90%. Three tonnes of PAX will be mixed per makeup batch, which will dilute the reagent to 10% (w/v) and will last for 25 hours of operation.

17.2.14.8Sulphuric Acid

Sulphuric acid will be used for pH adjustment prior to pyrite flotation, fed to the pyrite conditioning tank. It will be delivered to site as a liquid in 1 m3 IBC totes with a purity of 98%. Three totes will be kept in the dosing area, which will be enough supply for 15 hours of operation.

17.2.14.9Sodium Cyanide

Sodium cyanide will be used in lead flotation and will be fed to the SAG mill, lead conditioning tank, lead rougher flotation cells and to the lead cleaner flotation surge tank. It will be delivered to site as solid briquettes packed in one-tonne bulk bags in boxes. The briquettes will be mixed at a rate of one tonne per batch, diluted to 20% (w/v). This will be sufficient supply for 22 hours of operation.

17.2.14.10Lime (CaO)

Lime will be used in zinc flotation and will be dosed to the zinc conditioning tanks, zinc rougher flotation, zinc cleaner flotation surge tank, and the first, second and third zinc cleaners. It will be delivered to site as a quicklime powder. It will be stored in a 150-tonne silo and slaked for dosage. The lime slurry will be prepared at a 20% (w/w) solids concentration. The storage tank will have a 24-hour residence time, and the dosage system will have a ring main for constant recirculation to prevent scaling. Lime will be delivered to site in full truckloads pneumatically offloading into the silo, requiring approximately one truckload per day.

17.2.14.11Anti-Scalant

Anti-scalant will be used to prevent process water scaling and will be dosed to the process water pond pump suction line. It will be delivered to site as a liquid in 1 m3 IBC totes. One tote will service approximately 3.6 days of operation.

17.2.15Plant Services

17.2.15.1Air

Plant and instrument air will be supplied from air compressors. The air will be dried before it is distributed to various air receivers and users.

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17.2.15.2Water

Raw water will be sourced from the city sewage treatment effluent into a Raw Water Basin and transferred into a combined raw and fire water tank located in the process plant. Raw water will be used for gland service water, reagent mixing, dust suppression in the crushing circuit, process water makeup and the fire water reserve.

The main process water will be stored in a pond and will be supplied by thickener overflows, tailings management facility decant return, and raw water makeup. Separate lead and zinc process water tanks are fed from the lead and zinc concentrate thickeners respectively and used for the last cleaning and filter stages in each respective circuits to avoid contamination of the final products.

Potable water will be sourced from an aquifer and will be treated through an onsite treatment plant prior to use at the camp and in the operations facilities.

17.2.15.3Fuel

Plant mobile equipment will be powered by diesel, which will be stored onsite in a tanker and distributed to fueling stations.

17.3Plant Consumption

17.3.1Water

Nominal water consumption figures are provided in Table 17.3.1.

Table 17-2    Daily Water Consumption (Nominal)

Type			Daily Consumption (m3/d)
Raw Water			2,997
Potable Water			128

17.3.2Energy

A breakdown of the estimated annual energy consumption for the process plant is provided in Table 17-3.

Table 17-3    Annual Energy Consumption

Area			Annual Energy Consumption (MWh/y)
Feed Preparation – Crushing & Stockpiling			2,012
Milling – Grinding & Classification			93,487
Flotation & Concentrate Handling			36,512
Tails Thickening & Handling			2,701
Reagents			1,841
Water Services			2,692
Air Services			4,576
Fuel Storage & Distribution			102
Water Supply, Sewage, TSF, Plant Infrastructure (Buildings & Offices)			2,494
Total			172,396

17.3.3Reagents and Consumables

Annual consumption figures for each reagent and major consumables are shown in Table 17-4.

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Table 17-4    Annual Consumption of Reagents and Consumables

Reagent			Annual Consumption1 (t/y)
Flocculant			60
Zinc Sulphate			1,051
Aero 3418A			66
MIBC			310
Copper Sulphate			657
SIPX			95
PAX			876
Sulphuric Acid			2,920
Sodium Cyanide			350
Lime			6,862
Anti-Scalant			91
Grinding Media – SAG Mill			1,679
Grinding Media – Ball Mill			2,307
Grinding Media – Regrind Mills			59

1.Reagent consumptions are provided at supply concentration

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18.0PROJECT INFRASTRUCTURE

18.1Overall Site Development

The overall site plan is shown in Figure 18-1 and includes major facilities of the Project, including the proposed open pit mines, underground mine, crushing and feed preparation facilities, flotation process plant, electrical substations, administrative buildings, and tailings facilities.

18.2Site Access Road

18.2.1Access to Site

The project site is located approximately 450 kilometers from the capital of Marrakesh, with the main access being via the National 10 highway. The N10 passes through the town of Tinejdad, which is connected to the mine site through an upgraded dirt road.

The Port of Agadir, approximately 600 km away, will enable the transportation of goods and products for shipping, with an expected truck round trip from the mine to the port and back taking about 24 hours. The Nador Beni Ansar Port, also located approximately 600 km away, can also be utilized. All ports are accessible via national highways.

18.2.2Project Site Roads

Internal roads will provide access between the open pits, underground portals, crushing facilities, process plant area, mine service areas, and tailings storage facilities. These roads will be six metres wide and constructed flush with bulk earthworks pads to facilitate stormwater management.

18.3Process Plant Facilities

The infrastructure will support mining, processing, and construction activities. Located north-east of the central zone open pit, the main site will host a ROM pad, crushing and feed preparation facilities, stockpile with live reclaim, processing plant operations, and a security and site office buildings. Open pit mines will operate independently but connect to the central site via haul roads and mine access roads. A portion of the mining activity will also be carried out through an underground portal.

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Figure 18-1    Overall Site Plan

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Figure 18-2    Process Plant Layout

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Key infrastructure includes:

•Run of mine (ROM) pad.

•Crushing and feed preparation.

•Stockpile with reclaim chamber.

•Lead and zinc flotation plant.

•Lead and Zinc Concentration Handling

•Reagents Storage.

•Plant security and site office building.

•Plant administration and first aid building.

•Mine services area.

•Waste rock facility.

•Laboratory.

•Explosive magazine.

•Fuel and lubricant storage.

•Maintenance workshops and warehouse.

•Electrical Switchrooms.

Services provided cover power generation and distribution, water supply and distribution, surface water management, sewage and waste management, mine pit access, and haul roads. Portable / prefabricated structures will house security office and assay lab, while maintenance shops and warehousing will be brick and mortar buildings on concrete slabs.

Electrical Switchrooms will be located near areas of use.

Due to scarce water sources, an on-site water harvesting reservoir will be constructed near the process plant. A site water balance assessment will determine storage needs. A containerized water treatment system will ensure potable water availability for safety showers and personnel consumption.

Wastewater storage and treatment facilities will be established on-site. Solid waste, excluding domestic waste, will be managed off-site by licensed contractors, with on-site incineration for domestic waste. Dedicated areas for waste collection and sorting will facilitate disposal off-site, with no on-site long-term storage planned.

18.4Mining Facilities

18.4.1Truck Shop

A 1,000 m² truck shop will be adjacent to the ROM Pad for heavy equipment and light vehicles maintenance. There will be four bays dedicated to maintenance and one enclosed wash bay. All heavy equipment garage doors will be roll-type to maximize inside working space.

A 10-ton overhead crane equipped with two hooks will be installed in the heavy equipment maintenance area, while a single 5-ton overhead crane will cover the light vehicle area. Other maintenance equipment will be available for the workers, considering both the mining and the site maintenance fleets.

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Various sections will be installed in the truck shop, including a tool crib, an oil and grease deposit with a carousel distribution system, a mechanical / electrical room, storage rooms, a foreman’s office, toilets, locker facilities, a lunchroom, and a meeting room.

Water on the shop floor will be collected to an oil separation system with an integrated coalescent filter to ensure efficiency even if emulsification occurs. A sand pre-separator is included in the wash bay.

18.4.2Site Roads

The haulage roads will be 20 m wide based on the mine trucks selected and will connect the open pits, the underground mine portals, the mineralized material stockpiles, the ROM pad and the waste rock stockpiles. The service roads to access ventilation raises and other mine services will be 10 m wide.

18.4.3Low-grade Material Stockpile

The low-grade mineralized material stockpile will be located near the ROM pad to allow for the deposition of low-grade material for eventual blending or mineralized material storage, should the crushing or process plant be shut down. The stockpiled material will be rehandled by a loader and will have a capacity of 3.5 Mt (1.7 Mm³), with a planned height of 5 meters. A more detailed design will be developed during the next phase of the project, to possibly increase the height of the stockpile.

In the absence of geochemical characterization results, it is assumed that material may be potentially acid-generating and/or leachable. Based on this, it is assumed that the mineralized material stockpile will require a foundation consisting of a 2 mm thick HDPE geomembrane with a 300 mm sand protection layer on each side. This assumption will be revisited at the next phase of the project.

Additionally, the stockpile will be located at least 60 meters from the surrounding watercourses and at least 50 meters from the open pit rim.

18.4.4Rom Pad

Mineralized material exceeding the cut-off grade and characterized as low grade material, as described in section 16.7.3, will be trucked and stored on the mill ROM pad. The material will be separated based on the NSR-Pyrite value into two stockpiles – low- and high-grade pyrite. The ROM will be sufficiently sized for the estimated LOM stockpile volume requirements.

The feed to the concentrator will be blended from the two stockpiles, and material will be fed into the crushing plant with a front-end loader.

18.4.5Waste Rock Stockpiles

Two waste rock stockpiles with a combined capacity of 187 million cubic meters (Mm³), equivalent to 504 million tonnes (Mt), are planned. The North waste rock stockpile will have a capacity of 130 Mm³ and will be required at the start of mining operations; however, its southern section is scheduled for construction later during ongoing operations. The South waste rock stockpile will have a capacity of 57 Mm³ and will receive waste rock from both the South pits and a portion of the Central pit.

Because a large portion of the waste rock will be used for backfilling in the underground mine, only 6 million tonnes will be added to the waste rock stockpile throughout the underground mining operations.

The design of the waste rock stockpiles will adhere to all relevant mining industry standards, utilizing conservative parameters at this stage due to the absence of geotechnical studies and geochemical characterization results. It is currently assumed that the stored waste rock could be potentially acid-generating and/or leachable. In this optic, it is assumed that each pile will require a foundation consisting of a

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2 mm thick HDPE geomembrane with a 300 mm sand protection layer on both sides, or a similar design that would prevent potential contamination. This assumption will be revisited at the next phase of the project.

Additionally, each stockpile will be situated at least 60 meters from surrounding watercourses and 50 meters from the open pit rim.

The following design criteria were used for the waste rock stockpiles:

•Bench slope 2.5H: 1V

•Bench height: 10 m

•Berm width: 10 m

•Maximum height: 100 m

•Waste in-situ density: 2.75 t/m³

•Swelling factor: 1.36

18.5Tailings Storage Facility

Boumadine Global Mining SA has appointed Epoch Resources (Pty) Ltd (Epoch) to review and provide the Qualified Person’s signature for the update to the Pre-Economic Assessment (PEA) for a Flotation Tailings Storage Facility (TSF). A preliminary site selection study was completed in 2024 by an independent third-party Moroccan based consultancy, Groupement des Consultants et Ingénieurs de Maroc (GCIM) for two potential TSFs. The study design criteria have since been updated from two TSFs to only one Flotation TSF, with an updated storage capacity and tailings production rate based on the Life of Mine (LOM) production.

Boumadine Global Mining SA has appointed Epoch Resources (Pty) Ltd (Epoch) to review and provide the Qualified Person’s signature for the update to the Pre-Economic Assessment (PEA) for a Flotation Tailings Storage Facility. The history of the study is as follows:

•The study was initially completed in 2024 by an independent third-party Moroccan based consultancy, Groupement des Consultants et Ingénieurs de Maroc (GCIM) for two TSFs (Cyanide and Flotation).

•In 2025, the study design criteria have since been updated from two TSFs to only one Flotation TSF B2, with a doubled storage capacity and doubled tailings production rate for the same Life of Mine (LoM).

•In 2026, the study design criteria have since been updated for an increase in storage capacity of the one Flotation TSF B2, with an increased LoM.

A summary of the key design criteria for the TSF is summarized in Table 18-1 below.

Table 18-1    Design Criteria

Item			Design Criteria			Flotation TSF (TSF B2)			Flotation TSF (TSF B2) Update 2025			Flotation TSF (TSF B2) Update 2026			Source / Comment
1			Tailings Annual Deposition Rate (Mtpa)			0.86			1.65			1.64			Aya
2			Life of Mine Storage (Years)			11.0			11.3			14.5			Aya
3			Total tonnage stored (Million)			9.5			18.6			24.0			Aya
4			Lining Requirement			Lined									Best Practice
5			Construction Methodology			Full Containment Walls									GISTM compliance

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The conceptual design for the preferred TSF site comprises the components that affect the major costs at a high level. The tailings will be deposited from a slurry pipeline into a lined basin, behind earth-fill downstream wall raises. Drainage measures are included along the upstream wall slope and within the basin. Outside of the TSF basin, storm diversion channels along the perimeter of the TSF and an emergency basin located downstream of the TSF manage storm events, and a return water basin receives water pumped from a floating barge on the TSF pool.

The key design changes entail updating quantities according to the revised dimensions of the TSF to accommodate the additional tonnage. The construction phasing in the 2025 study had five lifts. An additional sixth lift was included to accommodate the additional tonnage required by the revised design criteria.

18.6Water Supply

18.6.1Raw Water Supply System

Due to water scarcity, off-site water sourcing has been evaluated. This could be done by using treated wastewater from nearby communities. Three raw water lines will provide up to 4230 m3/d to the site, from the towns of Tinejdad, Goulmima and Tinghir. An on-site water treatment plant will be used to ensure that the water quality is acceptable for the process. The raw water will be stored in the on-site water harvesting reservoir prior to being transferred to the open top raw water tank for distribution.

Figure 18-3    Raw Water Sourcing

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It is assumed that raw water will also be obtained from nearby boreholes and non-contact surface run-off water. A rainwater catchment pond will also harvest water for raw water usage and supplement in the flotation plant.

Raw water will be provided from the raw water tank for use as fire water. Feed to the potable water treatment plant and other users at the process plant will be sourced directly from boreholes and other non-contact water sources.

18.6.2Fire Water Supply System

Fire water will be piped to all main facilities via buried underground fire water ring mains around each of the facilities. The raw water tank live volume will hold a four-hour reserve of fire water equalling 288 m3. In addition, all buildings will be equipped with hose cabinets and supplemented with handheld fire extinguishers of two types - general purpose extinguishers for inside plant areas, and dry type extinguishers for inside electrical and control rooms.

18.6.3Potable Water Supply

The potable water treatment plant will be designed to local drinking water guidelines. The water treatment plant is expected to include multimedia filtration for reduction of turbidity, followed by ultraviolet disinfection for primary disinfection, and the addition of sodium hypochlorite for secondary disinfection. Treated potable water from the potable water treatment plant will be stored in the plant potable water tank and distributed via the plant potable water pump in a piping ring main to serve all potable water users in the process plant, heap leach, feed preparation and crushing facilities.

18.7Water Management

Studies and assessments required to complete a site wide water management will be done as part of the next phase of project.

18.8Power Supply

The Plant is estimated to have a Maximum Demand (running) of an estimated 26.3 MVA and a connected load of approximately 30.9 MVA. The main substation is sized at 40 MVA @ 65.75% load capacity.

Primary power to the Boumadine site will be provided by Office National de l'Electricité et de l'Eau Potable (ONEE) via a 220 kV high-voltage overhead transmission line. This newly constructed high-voltage overhead line will tee-off from the existing high-voltage overhead line that supplies power to the town of Tinejdad. The overhead line will connect to the main substation located north-east of the process plant. Power will be supplied to all areas of the site via the main 40 MVA substation.

18.9Sewage and Solid Waste Management

The project will utilize a modular sewage treatment plant based on a bio-oxidation process. Sewage and wastewater from various areas of site will be collected and transported to the plant via pump stations, strategically located to ensure efficient gravity flow. The plant will treat the water to the required environmental standards before being released into the environment. Solid waste will be periodically removed from the facility for off-site disposal.

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18.10Security

The Process Plant site area will be surrounded by fencing and access shall be through a main gate located at plant north direction. A ‘Main Gate House’ at the entrance shall provide station to any security personnel with a more security resources stationed at the ‘Plant Security and Changehouse’ Building.

18.11Accommodations

Effort of local hiring will be made given the proximity of the nearby town of Tinejdad and reduce the need for onsite accommodation.

Offsite accommodation shall include 200 prefabricated rooms. Senior staff and expatriate employees will be accommodated in rental units in the town of Tinejdad.

A construction/ permanent camp of 200-300 people capacity will be setup to accommodate any on-site accommodation needs.

18.12Administration and Plant Buildings

Administration and Plant Buildings area will be situated at the plant north entrance. Administration area will include the following:

•Main gate House, Emergency Services / Clinic Building, Plant Security and Changehouse, Admin Building, Admin Mess, Project Office, Laboratory, Warehouse.

Plant area buildings, which are accessible after passing by the Plant Security and Changehouse, will include the following:

•Plant Office, Plant Mess, Plant Workshop.

Three Ablutions facilities are located between the Administration and Plant Buildings. One Ablutions facility will also be located inside the Process Plant area close to the ROM Pad.

18.13Concentrate Transportation and Loadout

The three concentrate products will be shipped by road to the port of Nador-West, located 641 km from Boumadine. The zinc and pyrite concentrates will be shipped in bulk, while the lead concentrate will be shipped in big bags. The road haulage of concentrate will be contractor operated, with transportation along national highways.

A concentrate handling facility will be located near or at the port of Nador-West. This facility will be covered and allow for approximately 250,000 t of storage capacity over a surface area of 13,500 m2.

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19.0MARKET STUDIES AND CONTRACTS

19.1Market Studies

The commodities produced at Boumadine are gold, silver, zinc and lead, in the form of concentrates. The concentrates are sold under the terms of preliminary offers received and detailed in Section 19.2. As most of the value is contained in the gold and silver, AYA considered there were no need to carry out a proper market studies at this stage of the project.

19.2Commodity price projections

This study used the following prices for the different commodities:

•Gold: $3500/oz

•Silver: $50/oz

•Lead: $0.90/lbs

•Zinc: $1.37/lbs

Metals prices were kept constant for the entire life of mine plan of Boumadine.

19.3Contracts

19.3.1Concentrates sales

Aya Gold & Silver contacted several potential off-taker for the different concentrates that will be produced at Boumadine. Three different commodity traders have well received the request of interest sent by Aya Gold & Silver and have provided preliminary terms for off-take agreement for the purchase of the Lead-concentrate, Zinc-concentrate and Pyrite-concentrate of the Boumadine project. These terms are summarized as follows:

•Lead concentrate:

•Lead: payable at 95%, subject to a minimum deduction of 3 units

•Silver: payable at 95%, subject to a minimum deduction of 50 grams

•Gold: payable at 95%, subject to a minimum deduction of 1 gram

•Treatment charges: based on the annual Lead Asian Benchmark, with a $10/dmt in favor of the seller.

•Refining charges: for silver, based on annual benchmark, with $0.15 premium per Ag payable ounces in favor of the seller; for gold, $20 per gold payable ounces.

Penalty: for arsenic penalty, $2.5 for each 0.10% above 0.6% arsenic content in the lead concentrate

•Zinc concentrate:

•Zinc: payable at 85%, subject to a minimum deduction of 8 units

•Silver: deduction of 3 ounces, then payable at 70%

•Gold: deduction of 1 gram, then payable at 70%

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•Treatment charges: based on annual Zinc Asian Benchmark, with a $15/dmt in favor of the seller.

•Refining charges: for silver, based on annual benchmark, with $0.15 premium per Ag payable ounces in favor of the seller; for gold, $20 per gold payable ounces.

•Penalty: for Arsenic penalty, $2.5 for each 0.10% above 0.6% arsenic content in the zinc concentrate; for Cadmium penalty, $2.5 for each 0.1% above 0.3% cadmium content in the zinc concentrate

•Pyrite concentrate:

•Gold: if the gold content is 2g/t in the pyrite concentrate, payable of 77%; if gold content is 3g/t in the pyrite concentrate, payable of 78%; and if the gold content is 4g/t in the pyrite concentrate, payable of 79%.

•Silver: payable of 78% if silver content is above 30 g/t in the each tonne of pyrite concentrate,

•Penalty: If the As content is less than 2.0% in the concentrate, there is no Arsenic Penalty

No refining charges for gold or silver in the Pyrite concentrate

19.3.2Other contracts

There are no other material contracts or agreements in place as of the effective date of this Report. AYA Gold & Silver has not hedged nor committed any of its production to an off-take agreement with any of the commodity traders that have expressed their interest to the Boumadine project.

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20.0ENVIRONMENTAL STUDIES, PERMITS, AND SOCIAL
OR COMMUNITY IMPACTS

20.1Project Setting

The Project is located in the Drâa-Tafilalet region of western Morocco, with the planned mine site located approximately 19 km south of the nearest town of Tinejdad. The biophysical and social characteristics of the planned mining area are defined by its semi-arid climate and relatively rural and remote nature. Urban settlements are concentrated along main transport routes with villages located along valleys with access to water resources. Climatic conditions influence local ecosystems and land uses including agriculture, concentrated in areas of water availability.

20.2Environmental and Social Studies

Limited environmental studies have been completed in the past. As part of the environmental and social work currently being undertaken, an environmental baseline sampling program was initiated by BGM in September 2024 and a high-level screening for critical biodiversity habitat was undertaken by SLR in September 2024.

The baseline sampling, covering soils, air quality and water, was undertaken by AfriLabs, a Moroccan-based laboratory specializing in mineral and environmental analysis. Preliminary findings from the surveys and screening assessment reflect the following:

•Air Quality: Preliminary results indicate elevated particulate matter less than 10µm in diameter PM₁₀ and dust deposition levels, likely due to a combination of natural arid conditions and historical mining activities on site. Gaseous pollutants (NO2, SO2) were well below guideline thresholds.

•Water Resources: water resources are influenced by the semi-arid nature of the Project area. The Oued Akrouz, which crosses the area, is ephemeral but plays a key role in local irrigation feeding the Barrage d’Akrouz and acting as a pathway for aquifer recharge. Several wells are located along the valley of the Oued supplying local villages and farmers with domestic and irrigation water. Preliminary results indicate background water quality in the wells around the douars, , with some elevated concentrations of metals and metalloids reflecting the local geology and geochemistry. In the area of historical mining activities, water in the old shafts reflect the local mineralized geology with high acidity and elevated concentrations of some metals, metalloids, and major ion compared to the background groundwater quality from the irrigation well samples..

•Soils and Sediments: Background soil samples, i.e. soil samples which are distant from the area of historical mining activities on site, show some differences in chemical composition compared to the soil and sediment samples on site. Preliminary results of all samples indicate elevated levels of heavy metals and trace elements related to the local geology and geochemistry, and in some instances due to historical mining activities. While the survey included a preliminary review of the suitability of the soil for agriculture, environmental factors such as rainfall and evaporation / evapotranspiration, are probably the key limiting factors on soil use for agriculture in the local area.

•Biodiversity: most of the survey area includes a varying terrain of hills, small mountains and rocky outcrops all sparsely vegetated with a low density of plant cover. Preliminary flora and fauna surveys have identified the potential for some species of importance to occur on site. This will be confirmed through detailed seasonal surveys. The closest protected area to the Project is the Aferdou (El Kheng) Biological Reserve located 35km northeast of the site. While the Project site falls within an internationally recognised area, the Oasis du Sud Marocain Biosphere Reserve, the core zone of this reserve is located

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approximately 200km west of the Project site. Further engagement is required during the ESIA to better understand the biosphere reserve and any requirements for the ESIA and Project.

Additional baseline surveys have been completed during the course of 2026 covering social, biodiversity, soil, water, air quality and noise. The results will be used to inform the international ESIA process outlined below. Studies covering the topics of biodiversity, hydrology, hydrogeology, soils and sediments, air quality, noise, GHG, climate change, stakeholder engagement, labour and working conditions, community health and safety, and other social elements are ongoing.

20.3Permitting and Applicable Standards

The Project will require national permits to support the development of the mine and the requirement to carry out an Environmental Impact Assessment (EIA) under Moroccan Law. The EIA under Moroccan Law, prepared by NOVEC, was completed and submitted in July 2026 to the “Centre Régional d'Investissement” (CRI) (Regional Investment Center), the regional administrative body responsible for deciding on the environmental acceptability of the Project based on the EIA. The CRI review process, which includes a three-week public consultation period, is underway, and a decision on the environmental acceptability of the Project is expected before the end of 2026.

In addition to meeting national permitting requirements, the Project is being undertaken to meet international ESIA lender standards.

In this regard, applicable standards that will be applied to the ESIA preparation include:

•The existing legal framework in Morocco with respect to environmental protection, EIA processes, sustainable development, climate change, water management, soil protection, waste management, landfill management, oil management, energy efficiency, air quality, pollution prevention and control, biodiversity protection and conservation, cultural heritage protection, property expropriation, mining, labour and working conditions, and occupational health and safety, as well as any regulation that may influence the content of the ESMP measures either for construction or operation;

•The Lenders policies, including the 2024 EBRD Environmental and Social Policy and supporting E&S Requirements, the 2012 IFC Performance Standards, and the World Bank Group Environmental, Safety, and Health Guidelines (general and sector-specific guidelines), and supporting good practice E&S reference documents; and

•The relevant AYA corporate policies, sustainability framework and requirements.

20.4Key Environmental and Social Issues

20.4.1Potentially Affected Communities

Available preliminary information on potential project affected persons reflects a site with no permanent settlement identified within the planned mine footprint, within an area that, in general, is sparsely populated. The surrounding landscape is used for grazing, livestock movement and access to water and other natural resources. Villages are located mainly, along the valleys of the wadis where there is access to water resources and related livelihoods such as date and olive farming, and livestock. The closest settlements to the planned mine site are the downstream villages and farms located along the Oued Akrouz; the nearest is the village and farms of Bouydoud, about 1.5km to the northeast. Farming is primarily subsistence and/or supplies local markets. To the southwest and west, the villages and farms situated in the adjacent valley, along the Oued Taghoucht, lie approximately 7.5 km from the planned mine site. Similarly, associated farms are for

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subsistence and/or local markets. With the allocation of land plots and the establishment of the Taghoucht) Reservoir occupation and farming in this area may increase over time. Existing roads and tracks providing access to the planned mine site are used by local residents, farmers and herders. No villages or farms are located along the planned access road.

Nomadic and semi-nomadic herders seasonally use parts of the wider Project area for grazing, livestock passage, temporary camps and access to water points. The location and intensity of this use vary according to rainfall, pasture availability and water availability.

The planned mine footprint is located outside irrigated oasis areas, on land historically characterised by collective tenure and agro-pastoral use. The Project powerline and water-supply pipeline corridors extend across a broader landscape that include agricultural plots, grazing areas, livestock routes, access tracks, water points, wadis and areas subject to seasonal and/or temporary pastoral use. However, the precise tenure status and formal or customary land-use rights within the Project footprint remain to be confirmed. The land acquisition for the Project facilities is expected to be done through a long-term lease of the required area, with initial discussions underway with local authorities. A decision is expected before the end of 2026.

Social baseline surveys conducted during 2026 will be used to inform the social impact assessment undertaken as part of the ESIA.

20.4.2Environmental and Social Management

As part of the ESIA, potential impacts associated with the Project will be identified and assessed in a systematic approach. The impact assessment will inform a comprehensive Environmental and Social Management and Monitoring Plan to be developed for the Project. It will include:

•Mitigation measures for construction, operations and decommissioning

•Framework management plans covering biodiversity, erosion and sediments, waste, air and noise, water, community health and safety, community development and investment, supply chain, cultural heritage, rehabilitation and closure

•Stakeholder engagement plan including grievance management

•Monitoring plans for Project performance and the receiving environment (water, air, noise and biodiversity) including social aspects

Water management planning will continue to include the following elements:

•Characterization of the groundwater systems on site

•Environmental testwork to understand the geochemical composition and acid generating potential of the host rock and mineral residues

•Development of a site-wide climatic water balance to understand the water circuit and confirm water use, storage and discharge volume requirements.

•Establishment of an engineered stormwater runoff system catering for the diversion of clean water runoff around Project sites, separation of clean and potentially dirty water runoff on site, storage of potentially dirty water runoff in appropriately designed and engineered facilities, and collection and re-use of potentially dirty water runoff.

•Flood protection analysis and measures, if needed.

•Prioritizing water reuse and recycling.

•Water treatment for potable water supply, raw water supply, and discharge to the environment (if discharge is required).

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20.4.3Mine Closure

A conceptual mine closure plan governed by Aya’s Environmental Policy will be prepared as part of the ESIA.

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21.0CAPITAL AND OPERATING COSTS

21.1Capital Cost Estimate

21.1.1Summary

The capital cost estimates for the Boumadine Project PEA was compiled by Lycopodium and is presented here in summary format. The capital cost estimate reflects the Project scope as described in this report.

All costs are expressed in United States Dollars (US$) unless otherwise stated.The capital cost estimate was developed in Q3 2025 and escalated to Q3 2026 pricing and deemed to have an overall accuracy of +50%/-30%. The capital cost estimate conforms to AACE International (Association for the Advancement of Cost Engineering) Class 5 estimate standards.

The capital estimate for the Project is summarized in Table 21-1 by the major cost elements.

Table 21-1    Capital Estimate Summary

Capital Expenditures ($M)			Initial			Sustaining			Total
Direct Costs			299			507			806
Open Pit Mining			52			59			112
Underground Mining			-			407			407
Processing Plant			175			-			175
Shipping Infrastructure			11			-			11
Electrical Line			18			-			18
Raw Water Supply			33			-			33
Tailings Storage Facility			9			30			40
TSF Closure Costs			-			11			11
Indirect Costs			65			-			65
Subtotal			364			507			871
Contingency			99			-			99
Total			463			507			970

The following exchange rates have been used in the compilation of the estimate:

Table 21-2    Currency Exchange Rate

Currency			USD
AUD			0.65
USD			1
GBP			1.35
Euro			1.15
Rand			0.06
CAD			0.72

The capital cost estimate is broken down into mining capex, process plant and infrastructure capex, and tailings facility.

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21.1.2Mining Capex

All costs are reported in 2025 USD unless otherwise specified and conversion rates used as per Table 21-2. Escalation of 3% has been applied to estimates from prior years where required. Contingency is excluded from this estimate and will be applied separately.

Capital costs for mining activities include:

•General surface infrastructure – Roads, electrical distribution and other surface facilities.

•Open Pit facilities – Mineralized material and waste stockpiles

•Underground development – Lateral and vertical

•Underground mobile equipment – Acquisition and sustaining costs

•Underground mine infrastructure – Ventilation, electrical facilities, dewatering and facilities

Equipment costs for open-pit mobile equipment are included in the contractor mining cost. No equipment is forecast to be purchased by the owner. Other than the first year of mining, all waste mining costs for the open-pit operations are classified as an operating cost . A portion of these costs could eventually be reclassified as capital stripping at a later phase. First year mining costs have been capitalized.

A summary of the capital costs is shown in Figure 21-1 and Table 21-3.

Figure 21-1    Capital Cost Summary – Mining

Table 21-3    Capital Cost Estimate - Mining

Area			$M
UG Capital Development - Lateral			263.9
UG Capital Development - Vertical			36.9
UG Mobile Equipment - Acquisition			35.9
UG Mobile Equipment - Sustaining			22.8

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General Infrastructure			10.4
Open Pit Infrastructure			56.5
Underground Infrastructure			47.0
Capitalized OP Mining Costs (Y1)			44.8
Total			518.2

21.1.2.1General Surface Infrastructure

Surface facilities include roads, electrical distribution, and a truck shop. Costs are summarized in Table 2.

The plan includes 10 km of site roads and 2 km of service roads, with respective unit costs of $252,000/km and $84,000/km. Site roads are 20 m wide, while the service road linking the ventilation raise to the plant is 10 m wide. Electrical distribution will cover the total road length of 12 km at a unit cost of $95,000/km.

The truck shop has an estimated cost of $4.9 M. This includes three service bays as well as auxiliary bays, office space and a warehouse.

Table 21-4    General Facilities Cost

Area			$M
Site Road			2.5
Service Road			0.1
Electrical Distribution			2.9
Surface Facilities			4.9
Total			10.4

21.1.2.2Open Pit Facilities

Open pit mining will be performed by a contractor, who will provide and maintain all equipment and materials required.

Construction of mineralized material and waste stockpile facilities is included in the OP estimate. Stockpiles assume that geomembrane is required throughout; this requirement may be reduced at a later date as a result of further environmental testing. No clearing is required due to lack of vegetation.

Unit costs include surface preparation, foundations, and construction costs. Construction of waste stockpiles will be distributed over the mine life according to the production plan. Costs are summarized in Table 21-5.

Table 21-5    Open Pit Facilities Cost

Area			m2			$/m2			$M
Mineralised Material Stockpiles			370,000			13.00			4.8
Waste Stockpiles			4,920,509			10.51			51.7
Total									56.5

21.1.2.3Underground Development

All underground development is forecast to be completed by a contractor. An average unit cost of $2,500 was provided by Aya for this estimate, based on current contracts. This applies to ramp and lateral development and includes all labour, materials and equipment.

A vertical cost of $1,800/m was estimated for vertical development. This assumes a mix of 3-meter diameter raisebores and drop raises. Raisebores will be used in egress raises and raises from surface. Drop raises will be used in all other raises. An allowance of $1,900/m was added for SafeScape in required egress raises.

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Table 21-6 summarises the underground development costs.

Table 21-6    Underground Development Cost

Area			$M
Ramp			145.7
Lateral Development			118.2
Vertical Development			37.4
Total			301.3

21.1.2.4Underground Mobile Equipment

Underground mobile equipment is included for all production activities. Mine development is expected to be completed by the contractor who will provide all required equipment.

Recent quotes were provided by Aya for most of the mobile equipment costs. Quotes for similar recent projects were used to estimate costs for the remaining equipment.

Pickup trucks are assumed to be replaced after the useful life of 5 years. An overhaul of 50% of the initial acquisition cost was included for all other equipment at mid-life. Equipment exceeding useful life was replaced at initial cost.

Table 21-7 summarizes the parameters used to develop the mobile equipment estimate.

Table 21-7    Mobile Equipment

Activity			Quantity			Unit price			Life Years			Sustaining
Jumbo Drill			6			$931,000			12			$465,500
Production Drill			4			$931,000			12			$465,500
Mobile Compressors for Drill			4			$113,000			12			$56,500
Emulsion Charger			3			$373,000			8			$186,500
30-tonne Mine Truck			10			$720,000			9			$360,000
10-tonne LHD 10 t			12			$634,000			9			$317,000
Bolter			7			$817,000			12			$408,500
Transmixer			2			$343,000			9			$171,500
Scissor Lift			3			$188,000			8			$94,000
Boom Truck			5			$363,000			7			$181,500
Grader			1			$96,000			15			$48,000
Fuel Lube Truck			1			$287,000			8			$143,500
Water Truck			1			$287,000			8			$143,500
Pickup Truck			15			$44,000			5			REPLACE
Mine rescue Truck			1			$93,000			16			$46,500
Total - Mine Life						$35,896,000						$22,762,000

21.1.2.5Underground Mine Facilities

Underground facilities costs include egress, ventilation, electrical, dewatering, and rock passes. The estimate is summarised in Table 21-8.

Table 21-8    Underground Facilities Cost

Area			$M
Ventilation – Main Fans			31.4
Ventilation – Auxiliary Fans			1.2

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Ventilation – Mercaptan			0.1
Electrical Stations			14.3
Total			49.0

Main fans costs have been assumed based on existing quotes for similar projects in Europe and Canada, the cost includes the Electrical substation, the fan with dampers and accessories, mechanical and electrical installation, 15% of material cost to account for freight and civil works. In total there are nine different fan system, one for each zone, the total cost for each fan is shown in the Table 21-9 below.

Table 21-9    Main fans cost

Ventilation Zone			$M
North 1.1 & 1.2			1.9
North 2			3.3
North 3.1			1.3
North 3.2			1.3
Center			4
South 5			1.9
South 4.1			1.3
South 4.2			1.3
South 3			2.6
South 2.1			3.3
South 2.2			3.3
South 1.1			3
South 1.2			3
Total			31.4

Auxiliary fans cost has been estimated based assuming a total of 27 fans are required initially at the start of the project, 22 for the production levels (two (2) per level) and five (5) for the ramp development. It then includes an allowance of 20% for spares per year.

Underground bulkheads with drop board regulators and double personnel doors for egress have not been included in the cost and need to be included in the next phase of the project.

A mercaptan system is included for all surface raises at a unit cost of $26,000, based on a quote for a similar project.

The estimated surface infrastructure costs are $2.9M, including the surface substation, LTE backbone and diesel generator sets. Underground infrastructure is estimated at $14.3M, including underground substations and distribution of power and communications.

21.1.3Process Plant and Infrastructure Capex

The Table 21-10 below shows the Capital cost breakdown of the Process Plant and Infrastructure.

Table 21-10    Process Plant and Infrastructure Capital Cost Breakdown

Process Plant Main Areas			Supply Cost			Installation Hours			Labour Installation Cost			Freight Cost			Subtotal Cost			Project Contingency			Project Total
000 Construction Distributables			3,886,500			5,000			198,620			-			4,085,120			1,021,280			5,106,400

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100 Treatment Plant Cost			89,252,329			1,243,741			31,605,949			12,470,446			133,328,724			36,760,703			170,089,427
200 Reagents & Plant Services			11,506,610			138,552			3,680,704			1,595,996			16,783,311			4,791,217			21,574,527
300 Infrastructure			10,063,976			81,544			1,132,876			1,249,289			12,446,141			4,584,031			17,030,173
500 Management Costs			19,965,140			49,913			-			-			19,965,140			4,991,285			24,956,425
600 Owners Project Costs			41,327,536			397,295			-			603,028			41,930,564			10,482,641			52,413,205
Infrastructure
Shipping Infrastructure			-			-			-			-			10,870,368			5,435,184			16,305,552
Electrical Line			-			-			-			-			17,204,211			4,301,053			21,505,264
Raw Water Supply			-			-			-			-			30,426,944			7,606,736			38,570,522
Grand Total			176,002,091			1,916,045			36,618,149			15,918,760			290,416,546			80,954,015			385,929,069

21.1.3.1Estimating Methodology

Engineering lists, general arrangement drawings and a layout 2D model have been produced with sufficient detail to permit the assessment of the engineering quantities for concrete, steelwork, mechanical and electrical for the process plant and associated infrastructure. Quantities have also been benchmarked against similar flowsheet and throughput executed Lycopodium projects.

The unit rates and labour rates are based on historical in-house information or budgetary quoted and reflect current project market conditions. Budget pricing for major mechanical equipment was obtained from suitable, reputable suppliers and contractors.

21.1.3.2Quantity Development

The Project works were quantified to represent the defined scope of work and to enable the application of rates to determine costs.

Quantity information was derived from a combination of sources and categorized to reflect the maturity of design information as follows:

•Material Take Offs: detailed quantities derived from drawings and Naviswork model of Boumadine.

•Estimated: includes quantities derived from sketches or redline mark-ups of previous project drawings/data by estimating or similar projects.

•Factored: quantities derived from percentages applied as a factor based on experience or similar projects.

•Allowance: lump sum added to the cost estimate to account for costs that are certain to occur, but they cannot be identified with any accuracy.

The derivation of quantities is provided in Table 21-4 below, weighted by value of the direct permanent works (i.e. excluding temporary works, construction services, commissioning assistance, engineering costs, escalation and contingency).

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Table 21-11    Derivation of Quantities

Classification			MTO Prepared			Factored
Earthwork						1
Concrete						1
Structural Steel						1
Platework						1
Mechanical Equipment			1			0
Process Piping						1
Overland Piping						1
Electrical Bulks						1
Electrical Equipment						1
Instrumentation and Control						1
Buildings						1

The table below identifies the sourcing of costs included in the estimate.

Table 21-12    Supply Cost Source

Classification			Total Supply Cost			Estimated / Historical Pricing			Budgetary Quote
F Mechanical			62,871,233			0.48			0.52

21.1.3.3Installation Basis

This component represents the cost to install the plant equipment and bulk materials on site or to perform site activities. Installation costs are further divided between direct labour, equipment and construction indirect costs.

The labour component reflects the cost of the direct workforce required to construct the Project scope. The labour cost is the product of the estimated work hours spent on site multiplied by the cost of labour, inclusive of overtime premiums, statutory overheads, and payroll burden.

The equipment component reflects the cost of the construction equipment and running costs required to construct the Project. The equipment cost also includes cranes, vehicles, small tools, consumables, and Personal Protective Equipment (PPE).

The installation of the mechanical equipment has been built up using historical data of similar projects. The other disciplines installation costs were estimated as a percentage of mechanical costs and validated against historical data from similar projects.

21.1.3.4Field Indirects

Construction indirect costs encompass the remaining cost of installation and include items such as offsite management, onsite staff and supervision above trade level, crane drivers, equipment and labour mobilization and demobilization.

Construction indirect costs for all direct labour is included for all works in the capital estimate. This is inclusive of PPE, travel and clothing. Earthworks rates are inclusive of fuel, maintenance and running costs of machinery. Construction equipment and project cranes are included in the capital cost estimate.

Field In-directs costs have been calculated based on a percentage of direct installation cost. The percentage applied is derived from similar projects.

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21.1.3.5Engineering Services

The estimate for engineering design and construction management services was factored based on previous Lycopodium experience for similar sized projects and based on an allowance typical for projects of similar size and nature.

21.1.3.6Vendor Representatives

Some equipment will require vendor representation during construction and or commissioning. A provision has been included in the estimate to cover the vendor representatives’ services, including mob/demob and site expenses. It is based on an allowance typical for projects of similar size and nature.

21.1.3.7Freight

The freight estimate was derived from the following:

•By calculation as a percentage of supply costs for items where freight tonnes and bulk volumes are yet to be determined.

•By benchmarking against similar projects.

21.1.3.8Owners Costs

A summary of the Owners costs is given below based on estimates. A total cost, including contingency is shown to be $55.03M. This includes some general costs, Plant and Admin Pre-Production, Spare Parts and Plant Mobile Equipment. Owner's costs were developed by Aya, and include costs for engineering studies, project-specific home office costs, field staffing, pre-production in-country expenses, and a construction camp.

Table 21-13    Owners Project Cost Total

Plant Area			Supply Cost			Installation Hours			Labour Installation Cost			Freight Cost			Subtotal Cost			Project Contingency			Project Total
610 Owners Costs - General			22,650,775			113,254			-			-			22,650,775			5,662,694			28,313,469
620 Plant & Admin Pre-Production			11,751,715			303,906			-			-			11,751,715			2,937,929			14,689,644
640 Spare Parts			6,037,500			-			-			-			6,037,500			1,509,375			7,546,875
670 Plant Mobile Equipment			2,953,922			-			-			633,180			3,587,102			896,775			4,483,877
Total			43,393,912			417,160			-			633,180			44,027,092			11,006,773			55,033,865

21.1.4Tailings Storage Facility Capex

The high-level cost estimate was determined for the Capital Expenditure (CAPEX), Engineering Fees, Operating Expenditure (OPEX) and Closure CAPEX costs, as described below and detailed in Appendix A. The total LoM cost estimate changed from $48.2 to 56.6M (458M to 538M MAD (Moroccan Dirhams)), and comprises the following:

•The total LoM CAPEX with contingency is $37.4M (355 MAD),

•The total OPEX associated with the TSF has been estimated at $6.8M US (64,7M MAD) over the 14.5-year LoM, based on typical full containment operational costs sourced from tailings dam management companies.

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•Engineering consultant and construction supervision fees for the TSF over the LoM amounts to $1.5M (14,0M MAD), and

•The rehabilitation and closure of the TSF and post-closure monitoring costs amounts to $11M US (104,1M MAD).

Table 21-14    LoM Capital Costs (Million MAD)

LoM Costs			TSF ($M)
TSF Capex			32.9
Contingency			4.5
Total CAPEX			37.4
Opex			6.8
Engineering Consultant Fees: Monitoring and EOR			1.5
Closure CAPEX			11.0
Total			56.6

21.1.5TSF Closing Costs

The rehabilitation and closure and post-closure monitoring costs were estimated to be $11.0 M (104,1M MAD) for the study, and includes the following:

•The closure design work amounts to $0.13M (1.2M MAD).

•Closure CAPEX is $10.6M (100.5M MAD) phased over three years. The assumptions of the CAPEX estimates are based on the following items and associated works.

•Spillway channel, and

•TSF basin capping layer.

•The construction supervision of closure works amounts to $0.25M (2,4M MAD).

21.1.6Sustaining and Closing Costs

The sustaining and closing costs are detailed in each respective section.

The overall capital costs for the project, including sustaining and closing costs, are shown in Table 21-15.

Table 21-15    Capital Expenditures including Sustaining and Closing Costs

Capital Expenditures ($M)			Initial			Sustaining			Total
Direct Costs			299			507			806
Open Pit Mining			52			59			112
Underground Mining			0			407			407
Processing Plant			175			0			175
Shipping Infrastructure			11			0			11
Electrical Line			18			0			18
Raw Water Supply			33			0			33
Tailings Storage Facility			9			30			40
TSF Closure Costs			0			11			11
Indirects Costs			65			0			65
Subtotal			364			507			871
Contingency			99			0			99
Total			463			507			970

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21.1.7Contingency

An amount of contingency has been provided in the estimate to cover anticipated variances between the specific items allowed in the estimate and the final total installed project cost. The contingency does not cover scope changes, etc., or the listed qualifications and exclusions.

Contingency has been applied to the estimate as a percentage allowance and is varied for different cost elements but typically around 25% for all areas except electrical costs where the contingency applied was 50%. The average project contingency is 27%. It should be noted that contingency is not a function of the specified estimate accuracy which should be measured against the project total that includes contingency.

21.1.8Qualifications and Assumptions

The capital estimate is qualified by the following assumptions:

The base date for the bulk of pricing for the estimate is Q3 2026

Prices of materials and equipment with an imported content have been converted to US$ at the rates of exchange stated previously in this document. All pricing has been entered into the estimate utilizing native currencies wherever possible.

Suitable construction/fill materials will be available from borrow pits within 2 km of the work fronts. It must be reviewed following the results of the geotechnical investigation.

There is no allowance for unforeseen blasting in the bulk earthworks cost estimates.

It has been assumed mobile equipment purchased and used by the owner’s construction team will be handed over to the owner’s operations team upon completion of construction. No allowance for additional mobile equipment has been made for operations.

21.1.9Exclusions

The following items are specifically excluded from the capital cost estimate:

•Permits and licenses

•Project sunk costs

•Federal and state taxes and duties

•Exchange rate variations

•Sustaining Capital Costs

•Fuel Distribution

21.1.10Process Plant Pre-Production and Working Capital Costs

The costs incurred by operations during the latter stages of construction and commissioning are included in the capital cost estimate but are derived in this estimate. Pre-production costs associated with mining are excluded.

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21.1.10.1Pre-Production Labour

The pre-production site administration expenses cost covers the establishment of operations during the eight months preceding start-up and includes provision for power consumption, mobile equipment, and other expenses during this period.

21.1.10.2First Fill Reagents and Opening Stocks

Costs have been allowed in the project financial model cash flow estimates to purchase the mill balls, regrind ceramic media and reagents needed to commission the plant. Other consumables and reagents required for the process plant first fill and opening stocks are expected to be paid for in the commissioning and ramp up period, from the operating cost budget.

Sufficient first fill reagents and consumables have been estimated for the initial ball mill steel ball load to fill the reagent tanks, fill the regrind mills with ceramic media, and for other plant consumable requirements. Opening stocks refer to the purchase of the reagents and consumables required to sustain the operations for 6 weeks, which is the minimum on-site start-up storage quantity nominated by Aya Gold and Silver.

Quantities allowed have been based on either consumption over the minimum period or minimum shipping quantities, considering package size.

21.1.10.3Vendor Representatives

This cost allows for specialist vendor representatives to oversee commissioning of their processing equipment and include allowances for labour, airfares, and expenses.

21.2Operating Cost Estimate

21.2.1Mining Opex

Operating costs for mining activities include:

•Open-pit variable costs –mineralized material and waste mining, grade control and haulage

•Open-pit fixed costs – owners’s labour

•Underground variable costs – definition drilling, developing, stoping, haulage and backfill

•Underground fixed costs – labour, maintenance, electricity, mine services

A summary of the operating costs is shown in Figure 2 and Table 7. Year 1 open pit mining costs have been capitalized.

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Figure 21-2    Operating Cost Summary – Mining

Table 21-15    Operating Cost Estimate - Mining

Area			$M
Open-Pit Variable Costs			1,066.9
Open-Pit Fixed Costs			14.6
Underground Variable Costs			414.9
Underground Fixed Costs			306.1
Total			1,802.6
Unit Cost			$44/t

21.2.1.1Open-Pit Costs

Open pit mining will be performed by a contractor. Unit costs were provided by Aya based on current contracts. The following costs were used.

Table 21-16    Open-Pit Unit Costs

Area			$/t
Mineralised Material Mining			1.50
Waste Mining			2.00
Grade Control			1.10
Surface Haulage			2.00

Owner’s labour was estimated on an annual basis based on the Aya salary tables. Senior managers are budgeted as expatriate labour while technical services are a blend of local and expatriate labour.

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Table 21-17    Open-Pit Owner’s Team

Area			QUANTITY
Captains			1
Mine Admin			1
Technical Services			20
Annual Labour Cost			$1.3 M

21.2.1.2Underground Variable Costs

Variable costs were calculated for all stoping activities, based on production cycles and provided consumables cost. Costs for definition drilling and development were provided as per current contracts. Cemented rockfill costs were benchmarked from a similar project.

Unit costs include all consumables (including a 10% loss allowance). Equipment operation and maintenance costs are included as a fixed cost. Drilling costs include electricity. As no provision is made for a compressed air network, mobile compressors are included for all production drills. Costs are shown in Table 21-18.

Table 21-18    Underground Variable Costs

Activity			$/t			$M
Definition Drilling			3			56.7
Operating Development			$2,000/m			217.2
Drilling			3.2			48.1
Blasting			2.69			40.4
Backfill – CRF			6.39			1.9
Surface Haulage			2			50.5
Total			21.96			414.9

21.2.1.3Underground Fixed Costs

Fixed costs include operating labour, equipment maintenance of the owner’s fleet, electricity and mine services.

LOM costs are summarized in Table 21-19.

Table 21-19    Underground Fixed Costs

Area			$M
Mining Labour			49.9
Maintenance - Parts and Diesel			87.6
Maintenance Labour			13.8
Staff and Technical Services			65.4
Electricity			82.4
Infrastructure Maintenance and Mine Services			6.9
Total			306.1

All labour is budgeted on an annual basis according to provided salary tables. All labour is local except for certain managers that are expatriate. 1 supervisor is allocated for every 8 employees. The labour estimate is shown in Figure 21-3.

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Figure 21-3    Underground Labour Profile

Maintenance costs were developed based on annual operating hours. Annual maintenance costs are based on a 20-30% factor of initial acquisition cost and then converted to hourly rates. Diesel is also factored based on operating hours. Mobile maintenance labour is estimated using a factor of 4 units per mechanic and 12 units per electrician. Twelve (12) mechanics and 12 electricians are included to maintain the mine infrastructure. Diesel is budgeted at $1.30/L. Mobile maintenance costs are summarized in Table 21-20. Development equipment costs are included in lateral development rates.

Table 21-20    Mobile Maintenance Costs

Activity			Annual Hours			Annual Mtc Factor			Parts $/hr			Diesel $/hr
Jumbo Drill			3,000			20%			-			-
Production Drill			3,000			20%			$62			$6
Mobile Compressors for Drill			3,000			20%			$8			$13
Emulsion Charger			1,500			20%			$50			$2
30-tonne Mine Truck			5,000			30%			$43			$30
10-tonne LHD 10 t			5,000			30%			$38			$19
Bolter			3,000			20%			-			-
Transmixer			1,500			20%			$46			$7
Scissor Lift			2,500			20%			$15			$3
Boom Truck			2,000			20%			$36			$5
Grader			2,000			20%			$10			$6
Fuel Lube Truck			2,000			20%			$29			$5
Water Truck			1,500			20%			$38			$4
Pickup Truck			2,000			20%			$4			$1
Mine rescue Truck			1,200			20%			$16			$1

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Electricity for production drills is included in the drilling unit cost. Electricity consumption for the mine is based on the following factors:

•4.85 kWh/T for dewatering

•1.37 kW/T for lighting and services

•~44,000 MWh/year for main fans

•~22,000 MWh/year for auxiliary fans

•Electricity is budgeted at $0.09/kWh for a total of $82.4 M.

Other fixed costs included in estimate include an annual allowance of $150,000 for infrastructure maintenance, $100,000 for mine services, and $326,000 for ventilation maintenance (1% of initial capital).

21.2.2Process Operating Cost Estimate

21.2.2.1Introduction

The operating costs have been compiled by Lycopodium based on costs developed by Aya Gold & Silver and Lycopodium for all processing related costs.

Operating costs for Mining and TSF are detailed separately in Section 21.2.1 and Section 21.2.3 Respectively.

The estimate is considered to have an accuracy of +50 / -30%, is presented in US$ and is based on information obtained or adjusted for the third quarter of 2026 (Q3 2026).

21.2.2.2Process Plant Operating Costs

Processing operating costs have been developed by Lycopodium for a life of mine (LOM) blend. It is expected that the plant will operate on a range of mineralized material blends. The LOM processing costs are a weighted average of the various mineralized material type processing costs based on the LOM blend.

Processing operating costs have been developed for a plant with an annual throughput equivalent to 2.92 Mtpa of fresh mineralized material plant feed at a P80 grind size of 58 µm for the SAB circuit, 16 µm for the Pb regrind circuit, and 15 µm for the Zn regrind circuit, based on a 24 hour per day operation, 365 days per year. The plant will operate at the following utilization:

•6,570 h/y (75%) - Primary jaw crusher with a crushed rock stockpile.

•8,000 h/y (91.3%) - SAB grinding circuit and all flotation, regrind, thickening, reagents, and services processing plant sections.

•7,008 h/y (80%) – Concentrate filtration and handling.

The operating costs have been compiled from a variety of sources, including the following:

•Labour pay rates and manning as advised by Aya Gold & Silver.

•Laboratory rates as advised by Aya Gold & Silver.

•Grid power costs as advised by Aya Gold & Silver at a rate of 0.090 US$/kWh.

•Consumable prices from supplier budget quotations obtained in Q3 2025. Aya Gold & Silver advice, and the Lycopodium database and escalated to Q3 2026.

•Crushing and grinding consumable prices from supplier budget quotations, OMC modelled crushing and grinding energy and consumables, based on physical mineralized material characteristics determined from comminution testwork and estimated material properties from OMC database.

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•Reagent consumptions, mass pulls and recoveries based on the results from a metallurgical testwork program.

•First principal estimates based on typical operating data / standard industry practice.

The processing operating cost estimate is summarized in Table 21-21. The relative proportions of each operating cost center for the processing and process plant (processing plus G&A) operating costs are shown in Figure 21-4.

The operating costs include all direct costs to allow production of concentrates. The battery limits for the processing operating costs are as follows:

•Mineralized material delivered to the ROM bin.

•Tailings discharged onto the tailings storage facility (TSF).

•Pb concentrate in bags, with Zinc and Pyrite concentrates on stockpiles.

Qualifications

The operating cost estimate presented in this section excludes the following:

•All sunk costs.

•ROM stockpile rehandling costs.

•General and Administration:

•All head office costs and corporate overheads.

•Withholding taxes and other taxes.

•Any impact from fluctuations in foreign exchange rate.

•Any escalation from the date of the estimate.

•Project finance costs.

•Interest charges.

•Corporate Overheads.

•Political risk insurance.

•GST and/or VAT.

•All costs associated with areas beyond the battery limits of the Study.

•Plant rehabilitation costs.

•Land or crop compensation costs (included in CAPEX).

•Subsidies to local communities.

•License fees.

•Royalties.

•No allowance for contingency.

•Any business interruption costs.

•Concentrate:

•Pyrite concentrate processing or further refining costs.

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•In transit insurance costs.

•Transport or transit costs.

•Process / Tailings:

•Tailings storage costs, including future lifts and rehabilitation.

•Tailings dust suppression costs.

•External Government required monitoring and compliance costs.

•Raw water sourcing costs.

•Environmental:

•Any rehabilitation or closure costs.

•Labour:

•Union fees.

•Contract labour other than maintenance (generally self-performed by Aya Gold and Silver).

•Overtime allowances.

•First fill/opening stocks are captured in the CAPEX

The operating cost estimate includes the following:

•Import duties on consumable unit costs (in the consumables cost).

•Costs for the preparation and assaying of mine grade control samples and routine laboratory tests on the site water samples (in the laboratory cost).

Figure 21-4    Processing Plant Operating Cost Breakdown

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Table 21-21    Process Plant Operating Cost Summary (US$, Q3 2025, +50 / -30%)

Process Stream Costs			LOM Blend
Cost Centre			US$/y			US$/t
Operating Consumables			27,350,727			9.37
Maintenance			5,153,388			1.76
Contract Maintenance			1,135,189			0.39
Power (Plant)			13,177,570			4.51
Concentrate Handling			1,686,485			0.58
Laboratory			407,880			0.14
Process Plant Labour			3,347,874			1.15
Total Process Plant			52,259,122			17.90
General & Administration
Labour (G&A)			3,530,125			1.21
Laboratory (Exploration, Mining, Environmental)			2,707,657			0.93
Miscellaneous Costs			9,252,233			3.17
Total G&A			15,490,014			5.30
Total Overall OPEX			67,749,127			23.20

21.2.2.3Power

The power cost estimate has been based on grid power at a unit cost of US$0.09/kWh. The average continuous power draw and power cost for the LOM blend by plant area is summarized in Table 21-22.

Table 21-22    LOM Blend - Power Cost Summary (US$, Q3 2025, +50 / -30%)

Plant Area			Connected Power kW			LOM Average Continuous Power Draw kW			LOM Power Cost US$/y
120 – Feed Preparation – Crushing and Stockpiling			437			230			181,043
130 – Milling – Grinding and Classification			14,774			10,672			8,413,850
150 – Flotation and Concentrate Handling			5,941			4,168			3,286,037
180 – Tails Thickening and Handling			446			308			243,132
210 – Reagents			350			210			165,695
220 – Water Services			423			307			242,274
240 – Air Services			776			522			411,866
250 – Fuel Storage and Distribution			16			12			9,173
330/350/360/370 – Water Supply, Sewage, TSF, and Plant Infrastructure			605			285			224,499
410 – Mining General1			7,414			5,171			N/A
Total			31,182			21,885			13,177,570

Note 1: Mining Power Costs are covered in Section 21.2.1

The power consumption for the crushers and ball mills has been calculated by OMC based on the physical mineralized material properties determined from comminution testing and database estimation. The power consumption for the remainder of the individual plant mechanical equipment items has been calculated from the load list derived from the vendor supplied equipment and estimated based on the installed motor size of individual items of equipment, excluding standby equipment, adjusted by efficiency, load, and utilization

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factors to arrive at the annual average power draw. This is then multiplied by total hours per annum and the electricity price to obtain the power cost.

21.2.2.4Operating Consumables

Costs for processing operating consumables, including reagents, liners, fuels, and process supplies have been estimated and are summarized for the LOM blend by plant area in Table 21-23.

Table 21-23    Consumables Cost Summary (US$, Q3 2025, +50 / -30%)

Plant Area			LOM Blend
			US$/y			US$/t
Crushing			769,499			0.26
Milling			9,904,499			3.39
Pb Regrinding			326,530			0.11
Pb Concentrate Handling			38,656			0.01
Zn Regrinding			281,370			0.10
Zn Concentrate Handling			33,395			0.01
Py Concentrate Handling			535,033			0.18
Reagents			13,457,236			4.61
Fuel (Plant)			2,003,970			0.69
Total			27,350,727			9.37

The consumables cost for mining is included in the mining operating cost.

The consumption of reagents and other consumables has been calculated from laboratory testwork and comminution circuit modelling at estimated material properties, calculated from first principles, or has been assumed based on experience with other operations. No additional allowance for process upset conditions and wastage of reagents has been made.

Reagent costs have been sourced from a combination of budget quotations, Client supplied costs, and in-house data relating to similar projects in the region. Transport and freight to site and import duties and taxes have been added based on vendor information. In absence of vendor information, a cost of 17.5% is added for transport, customs, and duties.

A diesel price, delivered to site, of US$1.34 per litre has been used, as provided by the Client. Diesel will be used in plant mobile equipment and the diesel consumption is based on industry standard vehicle consumption rates and estimated equipment utilization.

Allowances have been made for water treatment reagents.

21.2.2.5Maintenance

The plant maintenance cost allowance has been factored from the capital supply cost using factors from the Lycopodium database and is summarized for the LOM blend.

The allowance covers mechanical spares and wear parts, but excludes crushing and grinding wear components, grinding media, liners, and general consumables which are allowed for in the consumables cost.

The maintenance cost excludes payroll maintenance labour which is included in the labour cost. No additional contract labour has been allowed for.

Allowances for plant mobile equipment and general maintenance expenses have been made.

The mobile equipment allowance is based on unit costs for maintenance of the light vehicles, portable generators, and other mobile equipment for the process plant.

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General maintenance expenses include specialist maintenance software, maintenance manuals, training costs, vendor visits, and control system maintenance and license fees.

21.2.2.6Labour

The labour rates, manning levels and rosters used to determine the labour operating cost estimate have been provided by Aya Gold & Silver.

The plant labour cost includes all labour costs associated with site-based administration, plant operations, laboratory, and maintenance personnel. The plant labour cost excludes all mining personnel (included in the Mining cost category) and all head office (included in the General & Administrative Overheads costs).

The estimate of the labour contingent has been based on a four shift operation (two shifts working 12 hours per day, tow rotation shifts), to provide continuous coverage for the plant operation with allowance for leave and absenteeism coverage. Provision has been made in the manning numbers to accommodate annual and sick leave requirements.

21.2.2.7Laboratory Costs

The laboratory cost has been provided by Aya Gold and Silver for a similar operation in the region. The laboratory cost includes, 21,000 mining production samples, 6,000 exploration samples, 4,451 processing plant samples, 177 met lab samples, and 80 environmental samples per month.

21.2.2.8Services and Utilities

Mobile Equipment

Plant mobile equipment requirements have been agreed with Aya Gold and Silver. Mobile equipment costs provide for the rental fees, fuel, and maintenance of the mobile equipment fleet (excluding the mining fleet and mining light vehicles). The purchase cost of this equipment has been included in the capital cost estimate.

The rental fees, fuel, and maintenance costs for the mobile equipment are included in the maintenance cost center.

Water Supply

As informed by Aya Gold & Silver, no direct water supply costs are incurred and allowance have been made related to the mechanical equipment maintenance and associated power consumption for pumping water, which are captured under the respective headings.

21.2.3Site General and Administration

The general and administration costs is based on Lycopodium's internal database, and from information and costs provided by Aya Gold and Silver. Costs were benchmarked with other similar operations by Lycopodium.

The general and administration expenses include the following ongoing operating expenses:

•Site office expenses including communications and communication maintenance, office equipment and supplies, computer supplies and software licenses.

•Insurance expenses covering industrial special risks, third party liability, motor vehicle and security. Labour associated insurances (medical, death and disability and workers liability insurances) are included in the labour costs.

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•Financial expenses including banking charges, legal fees, auditing costs and accounting consultants. Concentrate refining and royalties are excluded as they are deducted directly from revenue in a separate cost category.

•Government charges including permits and environmental inspection fees.

•Personnel expenses such as first aid and medical costs, safety supplies, travel and accommodation, expatriate travel, international expat recruiting / relocation costs, training, recreational and local facilities costs, professional memberships and subscriptions, and entertainment allowances. The allowances for expatriate travel (international and regional) and international expat recruiting / relocation include all personnel (mining as well as administration and process plant).

•Contract costs for personnel transport (in country charter flights and on site bussing), camp, catering and cleaning, environmental compliance testing, OH&S and other consultants. The camp catering and cleaning contract cost includes all personnel (mining as well as administration and process plant).

•Community relations expenses including general expenses, community projects and scholarships.

•Infrastructure operating costs

•Co-disposal facility operating costs

21.2.4Concentrate Shipping and Handling Costs

Ground concentrate shipping costs were provided by Aya, and were developed in cooperation with local Moroccan logistics company. These costs were developed based on the concentrate production from the mine plan, and developed from first principles.

The shipping costs include road transportation, handling at the port, and associated port fees. The total of these costs were estimated to be 479 MAD/t, or $52.22 /t.

Maritime shipping costs were estimated at $25/t based on costs for similar operations and publicly available data for bulk transportation of concentrates along similar shipping routes.

21.2.5TSF Opex

The OPEX of $6.8M (64.7M MAD) over the life of mine for the TSF comprises:

•$2.4M (22.7M MAD) over the life of mine for maintenance of the piping, TSF and equipment,

•$0.5M (5M MAD) over the life of mine for pumping of recovering water,

•$1.0M (10M MAD) over the life of mine for operational management, pipeline management and equipment maintenance, and

•$2.8M (27.0M MAD) over the life of mine for quarterly inspections, monitoring and quarterly reports by the design engineer.

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22.0ECONOMIC ANALYSIS

22.1Introduction

The preliminary economic assessment is preliminary in nature, it includes Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves, and there is no certainty that the preliminary economic assessment would be realized. Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability.

A preliminary economic analysis has been carried out for the Project using a cash flow model. The model is constructed using annual cash flows by taking into account annual processed tonnages and grades. The associated process recoveries, metal prices, operating costs and refining charges, royalties and capital expenditures (both initial and sustaining) were also taken into account. The financial model was completed on a 100% project basis and includes a 3% gross royalty to the ONHYM (Office National des Hydrocarbures et des Mines). As a general rule, the financial assessment of projects of this nature is carried out on a ‘100% equity’ basis, i.e. the debt and equity sources of capital funds are ignored. No provision is made for the effects of inflation. Results are given before and after taxation. All amounts in this chapter are presented in United States Dollars (US$) and unless otherwise stated, are referred to ‘$’.

The model reflects the base case and technical assumptions shown in the preceding chapters of this report. Discounting has been applied from the first year of mine construction (Year 0 onward).

According to Canadian securities law, the outcomes of the economic assessments mentioned in this section constitute forward-looking information. Results depend on inputs that could differ considerably from those predicted here due to known and unknowable risks, uncertainties, and other factors

22.2Fiscal and Economic Parameters

22.2.1Royalties and Duties

The economic analysis incorporates royalties due to the ONHYM. The Government of Morocco also collects various taxes and duties on the importation of fuels, supplies, equipment, and outside services as specified in the Mining Code.

The ONHYM is entitled to collect a 3% royalty on the total revenues from metal production.

22.2.2Taxes

Aya’s understanding of current Morocco tax regulations are applied to assess the tax liabilities. Taxation calculations are based on a 35% corporate tax rate as prescribed by Moroccan tax regulations. It also includes a mining tax rate of $0.32/t extracted. It is also anticipated that the project is eligible to receive a tax holiday period of 5 years.

Depreciation allowance uses a linear method based on life of mine (14 years) for initial investment as prescribed by Moroccan tax regulations. Investments made during the course of the operations are amortized on the remaining years of life.

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22.2.3Economic Parameters

All costs have been estimated in US Dollars. All capital and operating costs are provided in Q3 2026 money terms. Inflation rates have not been applied in the financial model as the evaluation has been carried out on a real terms constant money basis.

No price escalation has been applied to the model, and cost input parameters were considered fixed for the life-of-mine for the purposes of this financial valuation.

No salvage value has been considered within the economic model for the Project equipment or infrastructure such as buildings, earth works, or remediation works.

The financial valuation has been completed using base-case prices shown in Table 22-1 below for the various metals produced, with a sensitivity analysis provided for Gold and Silver.

Table 22-1    Base Case Prices and Sensitivity Ranges

Metal Prices			Units			Base Case Price			Sensitivity Range
Gold Price			$/oz			$3,500			$1704 - $5250
Silver Price			$/oz			$50			$24 - $75
Zinc Price			$/lb			$1.37			$0.90 - $1.71
Lead Price			$/lb			$0.90			$0.60 - $1.13

22.3Project Timing

Key assumptions with respect to project timing used in developing the financial model are:

•For the purposes of the financial valuation, project capital expenditure is spent over 2 years prior to production (Year 0). In the actual planned ramp up, it is estimated that 40% of the capital cost is likely to be incurred in Year -1 (2 years before production) of the implementation period, and the remaining 60% of the capital cost incurred in Year 0.

•For the financial valuation it has been assumed that the capital expenditure is incurred in full during the period prior to project commissioning and ramp-up.

•Commissioning, ramp-up and initial mining occurs during Year 1 of the schedule.

•Process plant commissioned and first material in the plant in Year 1 of the schedule.

•It is assumed that the project will reach 100% of name plate capacity within six (6) months of commissioning, in line with experience at other similar projects that have been constructed by Lycopodium elsewhere in Africa during the last few years.

22.4Financial Modelling

The financial assessment of the Project was carried out on a 100% equity basis, not accounting for potential sources of funding which may include debt. The technical parameters and key assumptions described elsewhere in this report are reflected in the financial model, with the LOM Project cash flow shown in Table 22-2.

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Table 22-2    Key PEA Assumptions and Economic Parameters

Parameter			Units			Value			Source
General
Tonnes per day processed			tpd			8,000			Aya
Process Plant Capacity Year 1			%			75%			Aya / Lycopodium
Process Plant Capacity Subsequent Years			%			100%			Aya / Lycopodium
Shipping Capacity Year 1			%			75%			Aya / Lycopodium
Shipping Capacity Subsequent Years			%			100%			Aya / Lycopodium
Process Plant CAPEX year 1 Split			%			40%			Aya
Metal Prices
Gold Price			$/oz			$3,500			Aya/Lycopodium
Silver Price			$/oz			$50			Aya/Lycopodium
Zinc Price			$/lb			$1.37			Aya/Lycopodium
Lead Price			$/lb			$0.90			Aya/Lycopodium
Other Expenses
ONHYM Royalties			%			3.00%			Aya
Aya Technical Assistance			%			2.75%			Aya
Mining Tax (per mineralized tonne)			$/t			0.32			Moroccan Tax Law
Income Taxes			%			35.0%			Moroccan Tax Law
Payables
Lead Concentrate
Pb Payable			%			95			Prel. off-taker terms
Pb Minimum Deductible			%			3			Prel. off-taker terms
Au Payable			%			95			Prel. off-taker terms
Au Minimum Deductible			g/t			1			Prel. off-taker terms
Ag Payable			%			95			Prel. off-taker terms
Ag Minimum Deductible			g/t			50			Prel. off-taker terms
Zinc Concentrate
Zn Payable			%			85			Prel. off-taker terms
Zn Minimum Deductible			%			8			Prel. off-taker terms
Au Deduction			g/t			1			Prel. off-taker terms
Au Payable (After Deduction)			%			70			Prel. off-taker terms
Ag Deduction			g/t			93.3			Prel. off-taker terms
Ag Payable (After Deduction)			%			70			Prel. off-taker terms
Pyrite Concentrate
Au Payable Level 1
Au Minimum Grade			g/t			2			Prel. off-taker terms
Au Payable			%			77			Prel. off-taker terms
Au Payable Level 2
Au Minimum Grade			g/t			3			Prel. off-taker terms
Au Payable			%			78			Prel. off-taker terms
Au Payable Level 3
Au Minimum Grade			g/t			4			Prel. off-taker terms
Au Payable			%			79			Prel. off-taker terms
Ag Payable
Ag Minimum Grade			g/t			30			Prel. off-taker terms
Au Payable			%			78			Prel. off-taker terms
Deductables

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Parameter			Units			Value			Source
Lead Concentrate
Treatment Charge			$/t			100			Estimated based on historical TC Values
Au Refining Charge			$/oz			20			Prel. off-taker terms
Ag Refining Charge			$/oz			1.5			Prel. off-taker terms
Zinc Concentrate
Treatment Charge			$/t			200			Estimated based on historical TC Values
Au Refining Charge			$/oz			20			Prel. off-taker terms
Ag Refining Charge			$/oz			1.5			Prel. off-taker terms
Penalties
Lead Concentrate
Arsenic Penalty Level 1
As grade threshold			%			0.6			Prel. off-taker terms
As penalty interval			%			0.1			Prel. off-taker terms
As penalty per interval			$			2.5			Prel. off-taker terms
Zinc Concentrate
Arsenic Penalty Level 1
As grade threshold			%			0.6			Prel. off-taker terms
As penalty interval			%			0.1			Prel. off-taker terms
As penalty per interval			$			2.5			Prel. off-taker terms
Cadmium Penalty Level 1
Cd grade threshold			%			0.3			Prel. off-taker terms
Cd penalty interval			%			0.1			Prel. off-taker terms
Cd penalty per interval			$			2.5			Prel. off-taker terms
Pyrite Concentrate
Arsenic Penalty Level 1
As grade threshold			%			1			Prel. off-taker terms
As penalty			$			0			Prel. off-taker terms

AYA GOLD & SILVER INC.            308

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Table 22-3    Cash Flow Model

Parameter			Unit			yr -1			yr 0			yr 1			yr 2			yr 3			yr 4			yr 5			yr 6			yr 7			yr 8			yr 9			yr 10			yr 11			yr 12			yr 13			yr 14			yr 15			yr 16			yr 17			yr 18			Total
Mining
Open Pit Mining
Mineralized Material			kt			-			1,333			2,504			1,811			2,258			3,426			1,978			2,383			853			1,208			2,726			1,874			-			-			-			-			-			-			-			-			22,354
Waste			kt			-			18,650			42,487			53,186			52,725			51,573			53,001			52,613			54,114			53,788			48,423			23,897			-			-			-			-			-			-			-			-			504,457
Strip Ratio						-			14.0			17.0			29.4			23.3			15.1			26.8			22.1			63.4			44.5			17.8			12.8			-			-			-			-			-			-			-			-			22.6
Underground Mining
Mineralized Material			kt			-			-			-			-			567			825			1,472			1,518			1,875			1,845			2,000			1,936			2,000			2,000			2,000			857			-			-			-			-			18,895
Waste to Surface			kt			-			-			-			249			335			496			411			515			505			605			714			663			690			690			496			5			-			-			-			-			6,373
Processing
Feed to Processing Plant			kt			-			-			2,190			2,920			2,920			2,920			2,920			2,920			2,920			2,920			2,920			2,920			2,920			2,920			2,920			2,920			1,099			-			-			-			41,249
Pb			%			-			-			0.59			0.81			1.10			0.99			0.80			0.67			0.55			0.44			0.45			0.36			0.41			0.31			0.36			0.36			0.30			-			-			-			0.58
Zn			%			-			-			2.41			2.24			2.34			1.56			1.76			1.65			1.43			1.17			1.05			0.76			1.04			0.72			0.76			0.81			0.68			-			-			-			1.37
Au			g/t			-			-			3.31			1.50			1.30			2.85			1.59			1.72			1.31			1.53			2.24			3.02			1.34			1.40			1.15			1.12			1.14			-			-			-			1.77
Ag			g/t			-			-			93			80			121			119			92			77			40			39			58			45			30			32			43			36			34			-			-			-			63
Lead Concentrate			kt			-			-			36			66			89			80			65			54			44			36			36			29			33			25			29			29			9			-			-			-			660
Pb Grade			%			-			-			29.6			29.6			29.6			29.6			29.6			29.6			29.6			29.6			29.6			29.6			29.6			29.6			29.6			29.6			29.6			-			-			-			29.6
Au Grade			g/t			-			-			48.0			15.8			10.1			24.8			17.0			22.2			20.5			29.7			43.0			72.3			27.7			39.3			27.4			26.5			32.4			-			-			-			26.2
Ag Grade			g/t			-			-			2,877			1,794			1,990			2,198			2,103			2,107			1,321			1,623			2,362			2,297			1,340			1,927			2,178			1,802			2,049			-			-			-			1,999
As Grade			%			-			-			2.2			1.0			0.6			0.9			1.5			2.2			2.2			2.0			2.3			3.1			2.0			2.9			2.0			1.6			1.7			-			-			-			1.7
Zinc Concentrate			kt			-			-			69			85			89			59			67			63			54			45			40			29			40			27			29			31			10			-			-			-			736
Zn Grade			%			-			-			57.4			57.4			57.4			57.4			57.4			57.4			57.4			57.4			57.4			57.4			57.4			57.4			57.4			57.4			57.4			-			-			-			57.4
Au Grade			g/t			-			-			1.1			0.5			0.4			1.4			0.7			0.8			0.7			1.0			1.6			3.0			1.0			1.5			1.2			1.1			1.3			-			-			-			1.0
Ag Grade			g/t			-			-			137			127			182			271			185			165			98			119			194			209			103			159			201			157			176			-			-			-			164
As Grade			%			-			-			0.1			0.1			0.1			0.2			0.2			0.2			0.2			0.2			0.3			0.4			0.2			0.3			0.3			0.2			0.2			-			-			-			0.2
Pyrite Concentrate			kt									970			1,041			1,026			1,111			1,026			1,038			866			847			1,048			1,148			846			1,001			883			894			369			-			-			-			14,113
Au Grade			g/t			-			-			5.3			3.0			2.6			5.3			3.2			3.5			3.1			3.8			4.4			5.5			3.3			2.9			2.7			2.6			2.4			-			-			-			3.7
Ag Grade			g/t			-			-			87			93			142			130			109			90			55			56			67			48			43			39			59			49			42			-			-			-			77
As grade			%			-			-			1.5			1.2			1.0			1.2			1.7			2.1			2.0			1.5			1.5			1.5			1.5			1.3			1.2			1.0			0.8			-			-			-			1.4
S Grade			%			-			-			44.3			44.3			44.3			44.3			44.3			44.3			44.3			44.3			44.3			44.3			44.3			44.3			44.3			44.3			44.3			-			-			-			44.3
Metal Profile
Metal Produced
Pb			kt			-			-			11			19			26			24			19			16			13			11			11			9			10			7			9			9			3			-			-			-			195
Pb - Pb Concentrate			kt			-			-			11			19			26			24			19			16			13			11			11			9			10			7			9			9			3			-			-			-			195
Zn			kt			-			-			39			49			51			34			38			36			31			26			23			17			23			16			17			18			6			-			-			-			422
Zn - Zn Concentrate			kt			-			-			39			49			51			34			38			36			31			26			23			17			23			16			17			18			6			-			-			-			422
Au			koz			-			-			224			135			118			257			143			156			118			138			202			273			120			127			104			101			39			-			-			-			2,254
Au - Pb Concentrate			koz			-			-			55			33			29			64			35			38			29			34			50			67			30			31			26			25			10			-			-			-			557
Au - Zn Concentrate			koz			-			-			2			1			1			3			1			2			1			1			2			3			1			1			1			1			0			-			-			-			23
Au - Py Concentrate			koz			-			-			166			100			87			191			106			116			87			103			150			203			89			94			77			75			29			-			-			-			1674
Ag			koz			-			-			6,341			7,266			10,913			10,795			8,369			6,978			3,594			3,572			5,240			4,099			2,755			2,936			3,908			3,246			1,156			-			-			-			81,167
Ag - Pb Concentrate			koz			-			-			3,315			3,799			5,705			5,644			4,375			3,648			1,879			1,868			2,739			2,143			1,440			1,535			2,043			1,697			604			-			-			-			42,436
Ag - Zn Concentrate			koz			-			-			303			347			521			515			399			333			172			170			250			196			131			140			187			155			55			-			-			-			3,873
Ag - Py Concentrate			koz			-			-			2,723			3,121			4,687			4,636			3,594			2,997			1,544			1,534			2,250			1,760			1,183			1,261			1,679			1,394			496			-			-			-			34,858
Au-eq1			koz			-			-			354			292			333			454			307			295			203			217			303			351			185			186			179			167			62			-			-			-			3,889
Ag-eq1			koz			-			-			24,806			20,460			23,284			31,790			21,483			20,681			14,236			15,224			21,177			24,547			12,952			13,033			12,517			11,722			4,308			-			-			-			272,219
Metal Payable

AYA GOLD & SILVER INC.            309

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Parameter			Unit			yr -1			yr 0			yr 1			yr 2			yr 3			yr 4			yr 5			yr 6			yr 7			yr 8			yr 9			yr 10			yr 11			yr 12			yr 13			yr 14			yr 15			yr 16			yr 17			yr 18			Total
Pb			kt			-			-			10			18			24			21			17			14			12			10			10			8			9			7			8			8			2			-			-			-			176
Pb - Pb Concentrate			kt			-			-			10			18			24			21			17			14			12			10			10			8			9			7			8			8			2			-			-			-			176
Zn			kt			-			-			69			85			89			59			67			63			54			45			40			29			40			27			29			31			10			-			-			-			736
Zn - Zn Concentrate			kt			-			-			69			85			89			59			67			63			54			45			40			29			40			27			29			31			10			-			-			-			736
Au			koz			-			-			184			110			93			212			116			127			96			113			166			225			98			102			84			82			31			-			-			-			1,839
Au - Pb Concentrate			koz			-			-			53			31			26			60			33			37			28			32			47			64			28			30			24			24			9			-			-			-			527
Au - Zn Concentrate			koz			-			-			0			0			0			1			0			0			0			0			1			1			0			0			0			0			0			-			-			-			3
Au - Py Concentrate			koz			-			-			131			78			67			151			83			90			68			80			118			160			70			72			59			58			22			-			-			-			1,309
Ag			koz			-			-			5,341			6,107			9,253			9,214			7,098			5,904			2,995			2,997			4,448			3,485			2,300			2,482			3,321			2,743			979			-			-			-			68,668
Ag - Pb Concentrate			koz			-			-			3,149			3,609			5,420			5,362			4,156			3,466			1,785			1,774			2,602			2,036			1,368			1,458			1,941			1,612			574			-			-			-			40,314
Ag - Zn Concentrate			koz			-			-			67			64			177			236			139			101			6			26			91			76			9			41			70			44			18			-			-			-			1,165
Ag - Py Concentrate			koz			-			-			2,124			2,434			3,655			3,616			2,803			2,337			1,204			1,197			1,755			1,373			923			983			1,309			1,087			387			-			-			-			27,189
Au-eq1			koz			-			-			294			243			277			380			256			246			168			180			252			292			153			153			148			138			51			-			-			-			3,229
Ag-eq1			koz			-			-			20,613			16,986			19,362			26,627			17,899			17,189			11,769			12,571			17,650			20,428			10,682			10,715			10,346			9,660			3,550			-			-			-			226,049
Revenue
Lead Concentrate			$M			-			-			350			312			392			504			344			316			200			213			305			332			178			183			190			173			63			-			-			-			4,055
Pb			$M			-			-			19			35			47			42			34			28			23			19			19			15			18			13			15			16			5			-			-			-			349
Au			$M			-			-			184			110			92			211			117			128			97			114			166			224			99			104			85			83			32			-			-			-			1,845
Ag			$M			-			-			158			181			271			268			208			173			89			89			130			102			68			73			97			81			29			-			-			-			2,016
Penalties			$M			-			-			-1			-1			0			-1			-1			-2			-2			-1			-2			-2			-1			-1			-1			-1			0			-			-			-			-17
Deductions			$M			-			-			-6			-6			-9			-9			-7			-6			-3			-3			-5			-4			-3			-3			-3			-3			-1			-			-			-			-71
TC			$M			-			-			-4			-7			-9			-8			-7			-5			-4			-4			-4			-3			-3			-3			-3			-3			-1			-			-			-			-66
Zinc Concentrate			$M			-			-			91			111			122			88			92			85			69			58			57			45			51			38			40			41			14			-			-			-			1,000
Zn			$M			-			-			101			125			131			87			99			93			80			66			59			43			58			40			42			45			14			-			-			-			1,084
Au			$M			-			-			0			0			0			2			0			0			0			0			2			5			0			1			0			0			0			-			-			-			11
Ag			$M			-			-			3			3			9			12			7			5			0			1			5			4			0			2			4			2			1			-			-			-			58
Penalties			$M			-			-			0			-1			-1			0			0			0			0			0			0			0			0			0			0			0			0			-			-			-			-4
Deductions			$M			-			-			0			0			0			0			0			0			0			0			0			0			0			0			0			0			0			-			-			-			-2
TC			$M			-			-			-14			-17			-18			-12			-13			-13			-11			-9			-8			-6			-8			-6			-6			-6			-2			-			-			-			-147
Pyrite Concentrate			$M			-			-			566			396			418			709			431			432			299			340			502			629			290			302			273			257			97			-			-			-			5,941
Au			$M			-			-			459			274			235			528			290			315			239			281			414			560			244			253			208			202			77			-			-			-			4,581
Ag			$M			-			-			106			122			183			181			140			117			60			60			88			69			46			49			66			54			19			-			-			-			1,360
Penalties			$M			-			-			0			0			0			0			0			0			0			0			0			0			0			0			0			0			0			-			-			-			0
Deductions			$M			-			-			0			0			0			0			0			0			0			0			0			0			0			0			0			0			0			-			-			-			0
TC			$M			-			-			0			0			0			0			0			0			0			0			0			0			0			0			0			0			0			-			-			-			0
Total Revenues			$M			-			-			1,006			818			932			1,301			866			833			568			611			864			1,006			519			523			504			470			173			-			-			-			10,995
Operating Expenditures
Open Pit Mining Cost			$M			-			-			98			116			117			120			117			118			114			115			108			60			0			0			0			0			0			-			-			-			1,082
UG Mining Cost			$M			-			-			5			8			36			43			55			61			69			72			74			69			76			67			63			25			0			-			-			-			721
Processing Charges			$M			-			-			42			52			52			52			52			52			52			52			52			52			52			52			52			52			20			-			-			-			741
G&A			$M			-			-			16			16			16			16			16			16			16			16			16			16			16			16			16			16			8			-			-			-			225
TSF			$M			-			-			0			0			0			0			0			0			0			0			0			0			0			0			0			0			0			-			-			-			7
Raw water cost			$M			-			-			1			1			1			1			1			1			1			1			1			1			1			1			1			1			1			-			-			-			14
Concentrate Shipping			$M			-			-			98			109			110			114			106			105			88			85			103			110			84			96			86			87			36			-			-			-			1,416
Total Costs			$M			-			-			259			302			332			346			346			353			340			340			354			308			229			232			218			182			64			-			-			-			4,205
ONHYM Royalties (on sales)			$M			-			0			30			25			28			39			26			25			17			18			26			30			16			16			15			14			5			-			-			-			330
Mining Taxes			$M			-			0			1			1			1			1			1			1			1			1			2			1			1			1			1			0			0			-			-			-			13

AYA GOLD & SILVER INC.            310

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Parameter			Unit			yr -1			yr 0			yr 1			yr 2			yr 3			yr 4			yr 5			yr 6			yr 7			yr 8			yr 9			yr 10			yr 11			yr 12			yr 13			yr 14			yr 15			yr 16			yr 17			yr 18			Total
Aya technical assistance			$M			-			0			28			23			26			36			24			23			16			17			24			28			14			14			14			13			5			-			-			-			302
Total of Expenses			$M			-			0			318			350			387			422			397			402			374			376			405			367			259			263			248			209			74			-			-			-			4,850
Taxes (Excluding Interest)
EBITDA			$M			-			-			688			469			545			879			470			431			194			235			459			639			259			260			256			261			100			-			-			-			6,145
Amortization			$M			-			-			47			50			53			57			61			66			71			76			83			92			102			116			127			127			0			-			-			-			-
Use of LCF			$M			-			-			0			0			0			0			0			0			0			0			0			0			0			0			0			0			0			-			-			-			-
Taxable income			$M			-			-			641			419			493			822			409			365			123			159			376			548			158			145			129			134			100			-			-			-			-
VAT			$M			-			-			47			8			6			2			0			1			-2			0			2			-8			-14			1			-3			-7			-21			-			-			-			-
Taxes payable			$M			-			-			47			8			6			2			0			129			41			56			134			184			41			51			43			40			14			-			-			-			783
Capital Expenditures
Initial Capital Expenditure			$M			162			301			-			-			-			-			-			-			-			-			-			-			-			-			-			-			-												446
Sustaining Capital Expenditure			$M			-			-			30			36			34			45			43			41			43			38			43			41			40			42			23			0			0			4			4			4			340
Total Capital Expenditure			$M			162			301			30			36			34			45			43			41			43			38			43			41			40			42			23			0			0			4			4			4			970
Cashflow Statement
Revenue			$M			0			0			1,006			818			932			1,301			866			833			568			611			864			1,006			519			523			504			470			173			0			0			0			10,995
Production Costs			$M			0			0			259			302			332			346			346			353			340			340			354			308			229			232			218			182			64			0			0			0			4,205
Royalties			$M			0			0			30			25			28			39			26			25			17			18			26			30			16			16			15			14			5			0			0			0			330
Mining Taxes			$M			0			0			1			1			1			1			1			1			1			1			2			1			1			1			1			0			0			0			0			0			13
Aya technical assistance			$M			0			0			28			23			26			36			24			23			16			17			24			28			14			14			14			13			5			0			0			0			302
Taxes PaId			$M			0			0			47			8			6			2			0			129			41			56			134			184			41			51			43			40			14			-11			0			0			783
Capex			$M			162			301			30			36			34			45			43			41			43			38			43			41			40			42			23			0			0			4			4			4			970
Pre-Tax Free Cash Flow			$M			-162			-301			686			456			537			870			450			413			167			214			441			626			233			233			248			274			104			-4			-4			-4			5,477
Pre-Tax Free Cash Flow - 5% Discount			$M			-162			-301			670			424			475			734			361			316			122			149			291			394			140			133			135			142			51			-2			-2			-2			4,066
Post-Tax Free Cash Flow			$M			-162			-301			640			448			532			868			450			284			126			159			307			443			192			182			205			234			91			7			-4			-4			4,694
Post-Tax Free Cash Flow - 5%			$M			-162			-301			624			417			471			732			361			217			92			110			203			278			115			104			111			121			45			4			-2			-2			3,537

1Au-Eq and Ag-Eq are calculated at metal prices of Au = $3500oz, Ag = $50/oz, Pb = $0.90/lb, Zn = $1.37/lb

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Figure 22-1    Boumadine Project - Annual Free Cash Flow Profile

22.5Results

22.5.1Key Economic Results

On a pre-tax basis, the Project has a Net Present Value (NPV) of $4,066 M at a discount rate of 5%, an Internal Rate of Return (IRR) of 97%, and a payback period of 0.7 years. On an post-tax basis, the NPV is $3,537 at a discount rate of 5%, the IRR is 93%, and the payback period is 0.7 years. The payback period corresponds to the time at which the cumulative cash flow becomes positive.

The key economic results of the PEA for the Boumadine Project are presented in the tables below.

Table 22-4    Key Economic Results

Project Economics			Units			Base Case
						Pre-tax			Post-tax
NPV5%			$B			4.1			3.5
IRR			%			0.97			0.93
Payback			Years			0.7			0.7
NPV: Capex1			-			8.8			7.6
Revenue LOM			$B			11.0			-
Avg. Annual Revenue			$M/y			765			-
EBITDA LOM			$B			6.1			-
Avg. Annual EBITDA			$M/y			427			-
Cumulative FCF LOM			$B			5.5			4.7
Avg. Annual FCF			$M/y			381			327

1.NPV:Capex ratio is the ratio of Net Present Value, discounted at 5%, to the initial capital expenditure.

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Table 22-5    Additional Economic Model Outputs

Item			Units			Year 1-5			LOM
General
Total Mine Life			years						14.4
Open Pit Strip Ratio1			-			21.1			22.6
Throughput Capacity			tpd			8,000			8,000
Total Tonnes Processed			Mt			13.9			41.2
Open-pit			Mt			11.9			22.4
Underground			Mt			2.0			18.9
Processed Grade
Au			g/t			2.05			1.77
Ag			g/t			101.6			63.5
Zn			%			2.04			1.37
Pb			%			0.87			0.58
Au-eq			g/t			4.11			3.10
Ag-eq			g/t			287			216
Recoveries
Au rec (%)			%			96.1%			96.1%
Ag Rec (%)			%			96.4%			96.4%
Zn rec (%)			%			74.7%			74.7%
Pb Rec (%)			%			82.0%			82.0%
Metal Production
Gold Production			koz			877			2,254
Silver Production			koz			43,683			81,167
Zinc Production			Mlbs			467			931
Lead Production			Mlbs			219			431
Gold-Equivalent Production (Au-eq)			koz			1,740			3,889
Silver-Equivalent Production (Ag-eq)			koz			121,823			272,219
Cost per tonne milled
Mining			$/t milled			51.47			43.70
Processing			$/t milled			18.08			17.96
G&A			$/t milled			5.58			5.45
Tailings, Environmental and Water Management Costs			$/t milled			0.47			0.50
Total On-site Operating Costs			$/t milled			75.60			67.61
Product shipping			$/t milled			38.63			34.33
Royalties			$/t milled			10.65			8.00
Mining Tax			$/t milled			0.34			0.32
Total Cash Cost			$/t milled			125.22			110.26
OP Sustaining Capital			$/t milled			3.08			1.44
UG Sustaining Capital			$/t milled			9.41			9.85
TSF Sustaining Capital			$/t milled			0.95			1.00
Total Costs including Sustaining			$/t milled			138.67			122.55
Operating Cost per Ounce
Total Cash Costs1			$/oz AuEq			998			1,169
Total AISC2			$/oz AuEq			1,105			1,300

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1.Cash costs include mine-site operating costs such as mining, processing, and direct site G&A, as well as product shipping, royalties and mining taxes.

2.AISC is calculated as the sum of treatment and refining charges, onsite operating costs, sustaining capital costs, and closure costs, divided by the quantity of ounces equivalent produced.

22.5.2Life of Mine Cash Flow Analysis

Figure 22-2 shows the cumulative and annual cash flow for the project at a base case metal prices.

The cash flows starting for years 13 through 15 relate to closure and environmental monitoring.

Figure 22-2    Annual Free Cash Flow

22.5.3Project Parameter Sensitivities

Sensitivity analysis was carried out to determine the effect of changes to input parameters on the base-case financial model. Each sensitivity analysis was performed independent of others.

An overall Projects Economic Sensitivity based on Gold and Silver Price is shown in Table 22-6 below. The objective was to show a range of possible prices for gold and silver to best show the range of economic viability.

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Table 22-6    Project Economics Sensitivity - Gold and Silver Price

Parameter			Units			Break-Even1			- 25%			Base Case			Spot Price2			Upside
Gold Price			$/oz			1,704			2,625			3,500			4,472			5,250
Silver Price			$/oz			24			38			50			66.85			75
Zinc Price			$/lb			0.90			1.03			1.37			1.77			1.71
Lead Price			$/lb			0.60			0.68			0.90			0.85			1.13
NPV5% Post-Tax			$M			0			1,808			3,537			5,525			6,764
IRR Post-Tax			%			-			57%			93%			128%			147%
LOM Revenue			$M			5,441			8,169			10,995			14,237			16,288
LOM EBITDA			$M			910			3,482			6,145			9,200			11,133
FCF-Unlevered Post-Tax			$M			89			2,467			4,694			7,253			8,860
Payback Period Post-Tax			Years			-			1.3			0.7			0.5			0.4
Capital Efficiency Ratio			-			0.0			3.9			7.6			11.9			14.6

1.Gold and silver prices at which NPV5% - post-tax is equal to $0M.

2.Assumed Spot Prices as of September 3, 2026

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22.5.3.1Gold and Silver Price Sensitivity Based Parameters

A series of sensitivity analyses were performed at different combinations of gold and silver prices. The ranges are shown in Table 22-6.

Table 22-7    Base Case Prices and Sensitivity Ranges

Metal Prices			Units			Base Case Price			Sensitivity Range
Gold Price			$/oz			$3,500			$2,750 - $4,250
Silver Price			$/oz			$50			$28 - $62
Zinc Price			$/lb			$1.37			-
Lead Price			$/lb			$0.90			-

The results are shown in Tables 22-7 to 22-10.

Table 22-8    Gold and Silver Price Based IRR - Pre-Tax

						Au Price
Ag Price						2,750			3,000			3,250			3,500			3,750			4,000			4,250
			38			70		%	76		%	81		%	87		%	92		%	97		%	102		%
			42			74		%	80		%	85		%	90		%	95		%	100		%	105		%
			44			76		%	81		%	87		%	92		%	97		%	102		%	107		%
			50			82		%	87		%	92		%	97		%	102		%	107		%	111		%
			54			85		%	90		%	95		%	100		%	105		%	110		%	115		%
			58			89		%	94		%	99		%	103		%	108		%	113		%	118		%
			62			92		%	97		%	102		%	107		%	111		%	116		%	120		%

Table 22-9    Gold and Silver Price Based IRR - Post-Tax

						Au Price
Ag Price						2,750			3,000			3,250			3,500			3,750			4,000			4,250
			38			66		%	72		%	77		%	83		%	88		%	93		%	98		%
			42			70		%	76		%	81		%	86		%	91		%	96		%	101		%
			44			72		%	77		%	83		%	88		%	93		%	98		%	103		%
			50			78		%	83		%	88		%	93		%	98		%	103		%	108		%
			54			81		%	86		%	91		%	96		%	101		%	106		%	111		%
			58			85		%	90		%	95		%	100		%	104		%	109		%	114		%
			62			88		%	93		%	98		%	103		%	107		%	112		%	117		%

Table 22-10    Gold and Silver Price Based NPV5% Pre-Tax in $M

						Au Price
Ag Price						2,750			3,000			3,250			3,500			3,750			4,000			4,250
			38			2,466			2,794			3,122			3,450			3,778			4,106			4,434
			42			2,671			2,999			3,327			3,655			3,983			4,311			4,640
			44			2,774			3,102			3,430			3,758			4,086			4,414			4,742
			50			3,082			3,410			3,738			4,066			4,394			4,722			5,050
			54			3,288			3,616			3,944			4,272			4,600			4,928			5,256
			58			3,493			3,821			4,149			4,477			4,805			5,133			5,461
			62			3,698			4,027			4,355			4,683			5,011			5,339			5,667

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Table 22-11    Gold and Silver Price Based NPV5% Post-Tax in $M

						Au Price
Ag Price						2,750			3,000			3,250			3,500			3,750			4,000			4,250
			38			2,194			2,463			2,732			3,001			3,270			3,539			3,808
			42			2,373			2,642			2,911			3,179			3,448			3,717			3,986
			44			2,462			2,731			3,000			3,269			3,538			3,807			4,076
			50			2,730			2,999			3,268			3,537			3,805			4,074			4,343
			54			2,908			3,177			3,446			3,715			3,984			4,253			4,522
			58			3,087			3,356			3,625			3,894			4,163			4,431			4,700
			62			3,265			3,534			3,803			4,072			4,341			4,610			4,879

22.5.3.2CAPEX and OPEX Sensitivity Based Parameters

Sensitivity analyses were also evaluated based on capital expenditures and operating expenses.

The results are shown in Tables 22-11 to 22-14.

Table 22-12    CAPEX and OPEX Based IRR - Pre-Tax

						CAPEX
						-30%			-20%			-10%			0%			10%			20%			30%
OPEX			-30%			145		%	130		%	118		%	108		%	100		%	92		%	86		%
			-20%			141		%	126		%	114		%	105		%	96		%	89		%	83		%
			-10%			136		%	122		%	110		%	101		%	93		%	86		%	80		%
			0%			131		%	118		%	106		%	97		%	89		%	82		%	76		%
			10%			127		%	113		%	102		%	93		%	85		%	79		%	73		%
			20%			122		%	108		%	98		%	89		%	81		%	75		%	69		%
			30%			117		%	104		%	93		%	85		%	77		%	71		%	65		%

Table 22-13    CAPEX and OPEX Based IRR - Post-Tax

						CAPEX
						-30%			-20%			-10%			0%			10%			20%			30%
OPEX			-30%			142		%	127		%	115		%	105		%	97		%	90		%	84		%
			-20%			137		%	122		%	111		%	101		%	93		%	86		%	80		%
			-10%			131		%	118		%	107		%	97		%	89		%	82		%	76		%
			0%			126		%	113		%	102		%	93		%	85		%	79		%	73		%
			10%			121		%	108		%	97		%	89		%	81		%	75		%	69		%
			20%			115		%	103		%	93		%	84		%	77		%	71		%	65		%
			30%			110		%	98		%	88		%	80		%	73		%	67		%	61		%

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Table 22-14    CAPEX and OPEX Based NPV5% Pre-Tax in $M

						CAPEX
OPEX						-30%			-20%			-10%			0%			10%			20%			30%
			-30%			5,244			5,161			5,077			4,994			4,910			4,827			4,743
			-20%			4,935			4,851			4,768			4,685			4,601			4,518			4,434
			-10%			4,626			4,542			4,459			4,375			4,292			4,209			4,125
			0%			4,317			4,233			4,150			4,066			3,983			3,899			3,816
			10%			4,008			3,924			3,841			3,757			3,674			3,590			3,507
			20%			3,698			3,615			3,531			3,448			3,365			3,281			3,198
			30%			3,389			3,306			3,222			3,139			3,055			2,972			2,889

Table 22-15    CAPEX and OPEX Based NPV5% Post-Tax in $M

						CAPEX
						-30%			-20%			-10%			0%			10%			20%			30%
OPEX			-30%			4,544			4,461			4,377			4,294			4,210			4,127			4,044
			-20%			4,292			4,208			4,125			4,041			3,958			3,875			3,791
			-10%			4,039			3,956			3,872			3,789			3,706			3,622			3,539
			0%			3,787			3,703			3,620			3,537			3,453			3,370			3,286
			10%			3,534			3,451			3,367			3,284			3,201			3,117			3,034
			20%			3,282			3,198			3,115			3,032			2,948			2,865			2,781
			30%			3,029			2,946			2,863			2,779			2,696			2,612			2,529

22.6Conclusion

An economic analysis of the mine schedule at the metal base prices shown earlier in Table 22-6 shows the project is financially viable. The project is highly sensitive to the metal price and recovery. However, the project remains economically viable in all sensitivity cases shown in this section.

On a pre-tax basis, the Project has a Net Present Value (NPV) of $4,066M at a discount rate of 5%, an Internal Rate of Return (IRR) of 97%, and a payback period of 0.7 years.

On a post-tax basis, the NPV is $3,537M at a discount rate of 5%, the IRR is 93%, and the payback period is 0.7 years.

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23.0ADJACENT PROPERTIES

The following information has been summarised from publicly available information as well as from Aya 2025 PEA.

Several vein occurrences containing Ba, Pb, Cu and Zn, which vary in size and thickness, are recognized regionally. Some of these veins are subject to artisanal exploitation by individuals or small Moroccan companies, as shown in Figure 23‑1. Regionally, the Boumadine Project is located near several mineral deposits and exploration properties that have been historically exploited, are currently in production, under development or subject to ongoing exploration.

The Zgounder Ag Mine, owned and operated by Aya, within the Moroccan Anti-Atlas region, comprises structurally controlled Ag mineralization, with Ag occurring principally as native Ag and Ag-bearing minerals within a structurally complex hydrothermal system. The current NI 43-101 Mineral Resource Estimate, effective 30 June 2025, reports approximately 18.94 Mt of Measured and Indicated Mineral Resources containing approximately 100.3 Moz Ag, together with 0.41 Mt of Inferred Mineral Resources containing approximately 4.5 Moz Ag for both open pit and underground Mineral Resources.

The Imiter Ag Mine, located on the northern side of the Saghro Massif approximately 80 km west of the Boumadine Property, is operated by Société Métallurgique d'Imiter (SMI), a subsidiary of the Managem Group. Imiter comprises structurally controlled, epithermal-style Ag mineralization hosted principally by vein structures.

Publicly available information reported by Managem in its 2024 Annual Financial Report indicates Measured and Indicated Mineral Resources of approximately 11.45 Mt at 407 g/t Ag, together with approximately 14.52 Mt at 98 g/t Ag of Inferred Mineral Resources. The deposit is mined by underground mining methods, however Managem is also developing the Imiter Centre open-pit project.

The Imiter Bis property, located approximately 5 km south of the Imiter Mine, is a separate exploration property comprising two mining licenses covering approximately 27.6 km² and is 100% owned by Aya. Exploration has identified an approximately 4 km-long, structurally controlled polymetallic vein system, typically between 1.5 and 2.0 m thick, but locally reaching up to 10 m, with mineralization comprising Au, Ag, Pb, Zn and Cu. Drilling undertaken during 2021 and 2022 confirmed Au and Ag mineralization across several zones; however, no Mineral Resource Estimate has currently been reported for the property.

The Azegour deposit, located in the Western High Atlas approximately 60 km south of Marrakech, is a historical Cu-Mo-W skarn deposit associated with the intrusion of the Permian Azegour granite into Lower Cambrian carbonate and volcano-sedimentary rocks. Mineralization is hosted principally within metasomatically altered carbonate rocks and comprises molybdenite, chalcopyrite and scheelite, with sulphide mineralization developed predominantly during the retrograde stages of skarn formation. Historical mining produced approximately 900,000 t of ore grading 1.4–2.8% Cu, 0.2–0.7% MoS₂ and approximately 0.35% WO₃. Mining ceased in 1971, although subsequent exploration was undertaken by BRPM. No current Mineral Resource Estimate prepared in accordance with CIM Definition Standards or NI 43-101 has been identified.

The deposits and exploration properties described above demonstrate the variety of precious- and base-metal mineralization present within the broader region, ranging from structurally controlled Ag and polymetallic vein systems at Imiter and Imiter Bis, sediment-hosted stratiform Cu-Ag mineralization at Tirzzit, to intrusion-related skarn mineralization at Azegour.

The Qualified Person has not independently verified the information relating to the adjacent and regional properties described above, and mineralization or Mineral Resources reported for these properties.

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Figure 23-1    Location of the Imiter Mine

Source: Aya 2025 PEA

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24.0OTHER RELEVANT DATA AND INFORMATION

24.1Project Execution Schedule

The Project Execution Schedule was developed to provide a high level overview of all activities required to complete the project. The project execution plan is summarized and presented in Figure 24-1.

Figure 24-1    Project Execution Schedule

Source: Lycopodium 2026

24.2Pyrite Treatment

While not included within this report, Aya has continued to work on the metallurgical development for a pyrite treatment circuit. A portion of the gold and silver included in Boumadine is refractory to direct cyanide leaching. The refractoriness to cyanidation is due to the high sulfur content in the pyrite concentrate. Boumadine’s gold is in the form of fine grains (below 15 µm) and mainly associated with pyrite. A pre-treatment is necessary to oxidize the sulfur of pyrite prior to cyanide leaching.

Internal and external trade-off studies have indicated that roasting is the best option for pyrite pre-treatment prior to conventional cyanidation. Roasting would allow for reduced tailings generation and lower operating costs while also producing sulfuric acid, for which there is a high demand within the Kingdom of Morocco.

Testwork is on-going to determine optimal conditions for operation of a pyrite roasting plant. A flowsheet has been developed, and engineering will continue to be progressed in parallel to the mine and concentrator

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designs. This will provide flexibility for Aya to decide whether it will sell all of the pyrite or if a portion can be strategically treated within the country to produce doré ingots and a sulfuric acid byproduct.

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25.0INTERPRETATION AND CONCLUSIONS

25.1Geology

The Boumadine Project is hosted within Neoproterozoic volcanic and volcano-sedimentary rocks of the Tiddiline Formation (TTF) and comprises a structurally controlled polymetallic hydrothermal system developed within a shallow volcanic environment. Mineralization is principally controlled by steeply dipping northwest- to north-northwest-trending fault corridors that are spatially associated with felsic subvolcanic intrusions and hydrothermal breccias. Hydrothermal alteration is extensive and comprises quartz-sericite-pyrite (phyllic) alteration within the felsic volcanic sequence and propylitic alteration within the underlying mafic volcanic rocks, reflecting a long-lived hydrothermal system that evolved during repeated episodes of fault movement and magmatism.

Mineralization comprises steeply dipping massive to semi-massive sulphide veins, hydrothermal breccias, replacement sulphides and pyrite stockwork developed within broad alteration halos. The highest-grade mineralization is typically associated with massive sulphide veins dominated by pyrite, sphalerite, galena, arsenopyrite and chalcopyrite, with surrounding zones of disseminated sulphides and quartz-carbonate veining. Recent drilling and geological studies have demonstrated that mineralization extends into both the felsic and mafic volcanic successions and have refined the structural interpretation of the deposit, providing increased confidence in the geological model and continuity of the principal mineralized corridors. Although the precise genetic classification of the deposit remains under investigation, the 2026 MRE is based on the observed geological, structural and mineralization characteristics and does not depend on any specific genetic model.

The Project is at an advanced stage of exploration and geological understanding, supported by drilling and sampling, geological mapping, orientated core, geophysical surveys, historical mine plans as well as previous geological studies. In addition, no significant issues in the sampling and analysis were identified with the analysis of the QAQC programs. These studies provided confidence to the local scale geometry of the mineralization and grade distributions in the resultant Mineral Resources and has allowed SRK to produce a robust geological model with over 158 individual domains modelled. However, there are areas of lower geological confidence, classified as Inferred Mineral Resources and there are some domains SRK was not able to classify due to limited data. SRK notes that there is a degree of uncertainty associated with depletion survey as-builts and that the most significant underground workings are currently inaccessible.

SRK has declared a Mineral Resource Statement for the Boumadine Project. The Mineral Resources are limited to material contained within the conceptualised open pit shell above the NSR reporting cut-off of 60 USD/t and, below the pit shell, to material contained within grade shells (following the removal of isolated shapes) based on a 110 USD/t NSR cut-off that satisfy a minimum 1.5 m mineable thickness. The Mineral Resources are reported within the Boumadine Mining licence on an in-situ, undiluted basis and amount to:

•Indicated Mineral Resources (open pit and underground) of some 8,582 kt, at a mean grade of 121.7 g/t Ag, 1.9 g/t Au, 0.09 % Cu, 1.01 % Pb and 2.19 % Zn; and

•Inferred Mineral Resources (open pit and underground) of some 45,365 kt, at a mean grade of 56.8 g/t Ag, 1.8 g/t Au, 0.08 % Cu, 0.58 % Pb and 1.33 % Zn.

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25.2Mining

The OP mine accounts for 54% of the mineralized material mined (22.4 Mt), at an average NSR value of US$232/t. OP mining extends over eleven years (2029 to 2039) from three pit areas (Central, South and North), with a peak total material movement of 55 Mt/y and a LOM stripping ratio of approximately 23:1. Mining will be carried out by a contractor using a conventional truck and shovel fleet, with 10 m benches near surface and 20 m double benches in rock. Mineralized material will be hauled to three surface stockpiles (Low Grade, LG Pyrite and HG Pyrite), which will be used to regulate plant feed according to operational needs and the overall project strategy.

The UG mine accounts for 46% of the mineralized material mined (18.9 Mt), at an average NSR value of US$147/t. UG development starts in 2031, and production runs from 2032 to 2043 from five surface portals. Mining will be carried out using the modified Avoca method, with stopes 20 m high (floor to floor) and 20 m long, and a minimum mining width of 2.0 m. Material will be hauled by 33 t trucks, via ramps, to the surface stockpiles.

Capital lateral and vertical development will be performed by a contractor, while sill development and all stoping activities will be performed by owner crews.

CCE Mining has identified the following risks associated with the mine plan:

•Pit depth and incremental stripping ratio: the selected pit shells carry high stripping ratios (approximately 24:1 for the Central Pit and 30:1 for the South Pit). With a fixed contractor cost, the incremental stripping ratio at the base of the pits does not currently exceed the breakeven with UG mining. However, that might change if the contractor costs are correlated to haulage distance.

•Historical workings: UG mining between 1986 and 1992 extracted approximately 261 kt from four levels. Until these workings are located and surveyed, uncharted voids within or beneath the Central and South pits and near UG infrastructure represent a safety and design risk.

•Geotechnical data: pit slope and stope design parameters are based on quick-logged exploration core and joint mapping. No dedicated geotechnical drilling or laboratory testing has been completed.

•Hydrogeology: no site-specific hydrogeological investigation has been completed, groundwater inflows have not been estimated, and the dewatering system has been sized from typical operating practice.

•Dependence on rock fill availability for the modified Avoca method: the method relies on a consistent supply of RF and CRF, which requires a robust system for managing waste material, whether generated underground or transported from surface.

•Critical pre-production period: the pre-production phase is essential to the overall success of the project, and a detailed, well-coordinated schedule for this period is crucial.

•Shared underground equipment fleet: sharing mobile equipment between five portal areas may lead to inefficiencies due to travel distances.

•Contractor capacity: the peak OP material movement of 55 Mt/y requires a large contractor fleet. Contractor availability and rates have not yet been confirmed by budget quotations.

CCE Mining has identified the following opportunities:

•Optimizing pit depth and the OP to UG transition could reduce waste movement and advance UG access to higher-grade zones.

•Increasing the UG level spacing beyond 20 m could reduce development requirements, subject to geotechnical confirmation.

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•Pit slope angles may be steepened once dedicated geotechnical data are available.

•A battery-electric UG fleet could reduce ventilation airflow requirements and diesel consumption.

•Optimizing the cut-off and stockpile strategy could improve the grade profile of plant feed in the early years.

25.2.1Mining Geotechnical

Project risks and opportunities identified during mining geotechnical evaluations are as follows:

Risks:

•Some major faults have been delineated in the Boumadine geological model, however these do not transect intended production areas. The spatial distribution of adverse geological features is an unknown that will have to be resolved to achieve FS level of study.

•The UG mining depths evaluated are shallower than those in the updated mine plan. Stress magnitudes are likely to be more adverse than those estimated in this study.

Opportunities:

•Reasonable conservatism has been applied to the overall pit slope angle recommendations. There may be opportunity to increase slope angles once additional geotechnical data is available.

•Due to data limitations, reasonably conservative global rock strength values to the design evaluations have been applied. Laboratory testing efforts will substantially improve confidence in material strength parameterization and help to resolve geological controls and spatial variability. Opportunity to increase stope sizing (or reduce dilution expectations) may be resolved from more detailed material strength data.

•In situ stress assumptions applied to stope design are somewhat conservative for the depths considered. Later phases of study should rely on numerical stress modelling to improve confidence in likely induced stress conditions. Opportunity to increase stope sizing (or reduce dilution expectations) may be resolved from more detailed estimates of in situ stresses.

•Longitudinal stoping provides operational flexibility to adjust stope panel lengths, thus there is an opportunity for optimizing design for local variances (good or bad) in ground conditions. This flexibility is particularly important where the deposit body is narrow as production performance in narrow zones is likely to be more sensitive to HW and FW dilution.

25.3Processing

Conclusions from the metallurgical and comminution testwork conducted on the Boumadine deposit include:

•The average BWi from two tests was used in the comminution circuit design (13.1 kWh/t), with the remainder of comminution design criteria estimated from OMC’s database.

•Locked cycle flotation tests with three circuits (Pb, Zn and Py) achieved acceptable recoveries with grades in the salable range required. The lead and zinc circuits performed best with regrind after rougher flotation, including three cleaning stages and one cleaner-scavenger stage.

•Sulfur content can be estimated based on the metals’ content in the material, proven via multilinear regression analysis showing a strong correlation between the laboratory and calculated values (R2 = 0.98). By assuming constant concentrate grades and recoveries, the LOM pyrite concentration production could be estimated based on the sulfur content equation for varying metals’ grades in the deposit.

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•The Boumadine mineralized material is highly refractory and resistant to direct cyanidation.

Based on the metallurgical testwork completed thus far, the following unit operations have been selected for the PEA flowsheet:

•Single stage primary crushing to reduce the ROM material size to a P80 of 136 mm.

•Crushed rock stockpile with a live capacity of approximately 8,115 tonnes, allowing for 24 hours of live storage and additional capacity for periods of primary crusher maintenance.

•SAB type grinding circuit consisting of a SAG mill and a ball mill to produce a ground product P80 of 58 microns (µm), with hydrocyclones for particle size classification.

•Lead rougher flotation, classification, regrinding and cleaner flotation to produce a lead concentrate of 29.6% grade. The lead concentrate will be thickened, filtered, and bagged prior to shipment.

•Zinc rougher flotation, classification, regrinding and cleaner flotation to produce a zinc concentrate of 57.4% grade. The zinc concentrate will be thickened and filtered for stockpiling prior to shipment.

•Pyrite rougher flotation to produce a pyrite concentrate with 4.2 Au g/t and 81.0 Ag g/t. The pyrite concentrate will be thickened and filtered for stockpiling prior to shipment.

•Tailings thickening and storage in a tailings management facility.

The process plant operating cost estimate was developed in Q3 2025 using data from database projects, studies, and previous operations by Lycopodium, and Aya Gold & Silver. The operating cost estimate conforms to preliminary economic assessment accuracy of +50%/-30%. The estimate covers general and administrative (G&A), and processing costs. The unit operating cost per tonne of material milled is estimated at US$22.44/t processed.

Lycopodium have identified the following risks for the project:

•Additional variability testing is required focussing on specific pit locations to verify material properties used for design parameters and recovery predictions for the deposit body correlates with the established mine plan.

Thickening and filtration tests should be performed to confirm design assumptions and sizing. The selected filtration technology should also conform to required transport moisture limits yet to be determined.

Filtration of ultrafine material can prove to be challenging possibly requiring the use of drying pads to further reduce product moisture content prior to shipment. The additional material handling is not currently costed and has the potential to impact final product quality if not performed well.

25.4Infrastructure

The Project is supported by a fully integrated site layout including open-pit mines, underground access, processing facilities, and support services.

Site access via the N10 highway through Tinejdad; internal haulage and service roads link pits, portals, ROM pad, process plant, and stockpiles.

Process plant facilities include ROM pad, crushing and feed preparation, flotation plant, concentrate handling, reagent storage, laboratories, workshops, fuel storage, and administration buildings.

Water supply relies on pipelines from nearby towns, boreholes, runoff capture, and an on-site water harvesting reservoir; containerized treatment systems ensure process and potable water quality.

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Mining facilities include a 1,000 m² truck shop, low-grade stockpile (3.5 Mt), ROM stockpiles, and two waste rock stockpiles (215 Mm³ total) designed with HDPE liners and conservative geotechnical assumptions.

Tailings management uses a single lined Flotation TSF with downstream raises, internal drainage, storm diversion channels, an emergency basin, and a return water system.

Power supply delivered through a new 220 kV line feeding a 40 MVA substation serving all site loads.

Additional infrastructure includes sewage treatment, security fencing, communications, explosives storage, and limited on-site accommodation supplemented by housing in Tinejdad.

Concentrate transport will be contractor-operated, shipping by road to Nador-West Port with a covered 250,000-t storage facility.

The Authors have identified the following risks associated with the project infrastructure:

The supply of water will rely on off-site sources as the area receives little rainfall annually. It will be important to include redundancy and storage capacity to reduce risks associated with water supply.

The transportation of pyrite will require well-coordinated logistics due to the large quantities of material to transport annually.

25.5Environmental and Social

The biophysical and social characteristics of the planned mining area are defined by its semi-arid climate and relatively rural and remote nature. Urban settlements are concentrated along main transport routes with villages located along valleys with access to water resources. Climatic conditions influence local ecosystems and land uses including agriculture, concentrated in areas of water availability.

Limited environmental studies have been completed in the past. As part of the environmental and social work currently being undertaken, an environmental baseline sampling program was initiated by BGM in September 2024 and a high-level screening for critical biodiversity habitat was undertaken by SLR in September 2024.

The baseline sampling, covering soils, air quality and water, was undertaken by AfriLabs, a Moroccan-based laboratory specializing in mineral and environmental analysis.

Additional baseline surveys have been completed during the course of 2026 covering social, biodiversity, soil, water, air quality and noise. The results are being processed to inform the international ESIA process outlined below. Studies covering the topics of biodiversity, hydrology, hydrogeology, soils and sediments, air quality, noise, GHG, climate change, stakeholder engagement, labour and working conditions, community health and safety, and other social elements are ongoing.

The Project will require national permits to support the development of the mine and the requirement to carry out an Environmental Impact Assessment (EIA) under Moroccan Law. The EIA under Moroccan Law, prepared by NOVEC, was completed and submitted in July 2026 to the “Centre Régional d'Investissement” (CRI) (Regional Investment Center). In addition to meeting national permitting requirements, the Project is being undertaken to meet international ESIA lender standards.

A conceptual mine closure plan governed by Aya’s Environmental Policy will be prepared as part of the ESIA.

25.6Capital Cost Estimate

The preliminary economics of the Boumadine Project can be assessed using the capital and operational cost estimates detailed in this PEA Update. The calculations are based on an open pit and underground mining

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operation concept, the development of a processing plant, infrastructure, and the Owner's expenses and provisions.

The capital cost estimate conforms to Class 5 guidelines for a preliminary economic assessment level estimate with a +50%/-30% accuracy according to the Association for the Advancement of Cost Engineering International (AACE International). The capital cost estimate was developed in Q3 2025 and escalated to Q3 2026, based on Lycopodium’s in-house database of projects and studies, experience from similar operations and inputs from WSP and Epoch Resources.

The total initial capital cost for the Boumadine Project is $463 M and the life-of-mine sustaining cost is $507 M.

25.7Operating Cost Estimate

The operating cost estimate was developed in Q3 2025 using data from projects, studies, and previous operations from Lycopodium’s internal database and escalated to Q3 2026, excluding mining, concentrate shipping and TSF operation which were developed by their respective consultants. The operating cost estimate conforms to preliminary economic assessment accuracy of +50%/-30%. The estimate covers the general and administrative (G&A), and mining and processing. The unit operating cost per tonne of material milled is $67.61/t.

25.8Economic Analysis

The economic analysis was performed assuming a 5% discount rate typical for silver-gold projects. Cash flows have been discounted to the start of construction, assuming that the project execution decision will be made, and major project financing will be carried out at this time.

The pre-tax NPV discounted at 5% is $4.1B, with a pre-tax IRR of 97% and payback period of 0.7 years. On an after-tax basis, the NPV discounted at 5% is $3.5B, with an after-tax IRR of 93%, and payback period of 0.7 years. Cumulative after-tax unlevered free cash flow totals $4.7B. Tax calculations are based on Aya’s understanding of current Moroccan tax regulations as of the effective date of this report.

Readers are cautioned that the PEA is preliminary in nature. It includes Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves and there is no certainty that the PEA will be realized. Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability.

A summary of the project economics is listed in Table 25-1, and after-tax free cash flow is shown graphically in Figure 25-1.

Table 25-1    Key Economic Results

Project Economics			Units			Base Case
						Pre-tax			Post-tax
NPV5%			$B			4.1			3.5
IRR			%			0.97			0.93
Payback			Years			0.7			0.7
NPV: Capex			-			8.6			7.6
Revenue LOM			$B			11.0			-
Avg. Annual Revenue			$M/y			765			-
EBITDA LOM			$B			6.1			-
Avg. Annual EBITDA			$M/y			427			-

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Cumulative FCF LOM			$B			5.5			4.7
Avg. Annual FCF			$M/y			381			327

Figure 25-1    Boumadine Project - Annual Free Cash Flow Profile

25.9Conclusion

The Mineral Resources currently estimated for the Boumadine Polymetallic Project consist of:

•An Indicated Mineral Resource of 8.6 Mt grading 121.7 g/t Ag, 1.92 g/t Au, 2.19% Zn and 1.01% Pb containing an estimated 33.6 Moz of Ag, 532 koz of Au, 188 kt of Zn and 87 kt of Pb;

•An Inferred Mineral Resource of 45.4 Mt grading 56.8 g/t Ag, 1.82 g/t Au, 1.33% Zn and 0.58% Pb containing an estimated 82.8 Moz of Ag, 2.7 Moz of Au, 604 kt of Zn and 265 kt of Pb;

•The Updated PEA provides a base case assessment for developing the Boumadine mineral resource by conventional open pit and underground mining methods, and a crushing-milling-flotation processing plant with a tailings dam.

•The Updated PEA economic analysis shows the Boumadine Polymetallic Project has an after-tax NPV5% of $3.5B, IRR of 93%, and a payback period of 0.7 years. The PEA economics support a decision to continue to advance the Project and carry out additional detailed studies, including a pre-feasibility study.

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26.0RECOMMENDATIONS

26.1Summary

The results of the Updtaed PEA are sufficiently positive to warrant advancing the Project to the next study phase. It is recommended that the Project advance from the current PEA stage into the subsequent study phases, including Pre-Feasibility work toward a full Feasibility Study. Table 26-1 summarizes the proposed budget to advance the Project through the Feasibility stage.

Table 26-1    Budget for Advancing Project to Next Stages

Area			$M, USD
MRE Reporting			0.3
Metallurgical Testwork			1.3
Processing Plant			1.4
Mining Consultant and Geotechnical			1.5
Tailings Storage Facility and Geotechnical			0.5
Infrastructure			0.2
Hydrogeology			1.0
Environmental Studies, Permitting and Social Considerations			0.7
Contingency			0.7
Total			7.5

26.2Geology

SRK considers there to be opportunities to further improve confidence in the geological interpretation, MRE, and supporting datasets for the Boumadine Project. SRK recommends that Aya consider the following:

In relation to drilling and geological understanding:

•All drillholes should be designed to intersect the interpreted mineralized wireframes as close to perpendicular as practicable, thereby maximising representivity and location accuracy.

•Continue the ongoing infill drilling program to improve geological and grade continuity that support the conversion of Inferred Mineral Resources to the Indicated category.

•Complete additional twinned diamond drillholes (DD) at representative locations across the deposit, targeting selected existing RC, MP and DD holes completed between 2018 and 2026. Twin holes should be selected to provide representative coverage of the principal mineralised domains, mineralisation styles and grade ranges, one of the target areas should be where different drilling and sampling methods overlap. The twinning programme should be used to assess drilling and sampling precision, identify any systematic bias between drilling methods, quantify short-scale geological and grade variability (including the nugget effect), and provide additional information to support the confidence assigned to the current MRE.

•Complete additional step-out diamond drilling to test the strike and depth extensions of the known massive sulphide domains, where this remains open.

•Aya should investigate whether whole rock geochemistry will help with the identification and standardisation of the lithologies encountered in the drilling.

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•Include total sulphur (S%), and where practicable sulphide/sulphate sulphur analyses, within future drilling programs to improve/confirm definition of the oxidation profile and weathering boundary.

In relation to sampling and core management:

•Record the sample weight for every RC drillhole interval (including rejects) and compare this with the theoretical sample mass based on the logged lithology to identify intervals with poor recovery that may require re-drilling.

•Photograph all available drill core and RC chip trays and incorporate these records into the Imago database to establish a comprehensive digital photographic archive.

•Include all drillcore photographs taken prior to 2024 into the Imago database.

•Continue to maintain and periodically review sampling and QA/QC procedures as exploration progresses.

In relation to density:

•Expand the density measurement program, particularly within the mineralized wireframes and transitional (weathered) mineralized zones, to improve the representativeness of the density dataset and reduce reliance on regression-derived density estimates.

•Collect density measurements from drillholes that currently contain no density information and consider obtaining additional measurements from archived drill core where appropriate.

•Consider the use of plastic film or paraffin wax coatings during density determinations to minimize moisture ingress for friable or fractured core.

•Undertake a representative moisture content sampling program using fresh drill core collected from across the Project to improve future mining and processing studies, particularly for the Feasibility Study.

In relation to QA/QC:

•Expand the duplicate QA/QC database to include the complete analytical suite and include the following assays obtained as a minimum, Cu, Pb and Zn prior to the next Mineral Resource update.

•Introduce pulp duplicate samples into the routine QA/QC program.

•Consider introducing a commercially certified blank material or undertake a round-robin program to better characterise the locally sourced blank material.

•Review duplicate records where assay values are missing or recorded as zero and update the database against the original laboratory certificates where appropriate.

•Investigate and correct anomalous CRM assay values, including the Zn CRM records noting as being in excess of 100%.

•Consider discontinuing OREAS 100A, OREAS 354 and OREAS 994 as these are not certified ted for base metal analysis using Aqua regia or for Au and Ag using fire assay analysis.

•Investigate the small number of CRM sample swaps/mislabels that remain within the QA/QC database, resulting either from the mislabelling of original drill core samples or from other QA/QC samples (including CRMS) being incorrectly assigned, and resolve these where possible.

•Continue the routine umpire laboratory program and consider increasing the proportion of samples submitted to the umpire laboratory, including representative samples from the 2024 drilling program.

•Where practicable, ensure that future umpire analyzes use analytical methods with reporting limits comparable to those of the primary laboratory.

•In relation to surveying

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•Incorporate the planned high-resolution pre-mining drone survey into future geological models and associated mining studies.

•Where safe and practicable, resurvey historical underground workings (once mining begins) and surface artisanal excavations to improve confidence in future depletion models.

In relation to Resource Estimation:

•Develop a 3D lithological model, subject to the availability of sufficient geological data, to improve geological understanding and support future Mineral Resource, Mineral Reserve, geotechnical, and mine planning studies.

•Undertake a change-of-support study following additional infill drilling to optimize the parent block dimensions and improve future Mineral Resource estimates.

•Incorporate a Mineable Stope Optimizer (MSO) into future Mineral Resource estimates using parameters representative of the proposed underground mining method for Resource reporting.

•Aya should expand its metallurgical test work program to include the weathered/transitional material in order to demonstrate this has reasonable prospects for eventual economic extraction.

In relation to Net Smelter return Calculation and Equivalent cut-off value calculations:

•Aya should review the NSR royalty calculation in future updates so that royalties are applied directly to the relevant concentrate streams and remain linked to the underlying value drivers.

•SRK recommends that Aya model concentrate penalties and deductions separately from treatment charges, using concentrate grades as inputs.

•Aya should consider incorporating freight costs within the NSR calculation in future updates, rather than operating costs, to link these costs directly to the salable concentrate mass.

•Review the concentrate mass calculation methodology to account for the sequential processing flowsheet. Specifically, Pb concentrate should be based on plant feed grades, Zn concentrate should be based on Pb-circuit tailings grades, and pyrite concentrate should be based on Zn-circuit tailings grades. This would affect the calculated mass pull and therefore shipment costs. The difference is expected to be immaterial for Pb and Zn concentrates but may be more relevant for pyrite concentrate.

•SRK recommends calculating the equivalent Ag and Au grades using NSR factors, opposed to the current approach which uses metal prices and metallurgical recoveries only. This approach will avoid any inconsistency with the NSR framework already in place.

SRK has not prepared a detailed exploration budget for the Project but has reviewed the budget and associated schedules provided by Aya. These principally relate to infill drilling intended to upgrade portions of the currently classified Inferred Mineral Resources to Indicated Mineral Resources and support the proposed 15 year life-of-mine plan. Based on Aya's proposed program, approximately 326 000 m of additional drilling is anticipated, with estimated drilling and sampling costs of approximately US$65 million. This cost estimate excludes the multidisciplinary Feasibility Study (FS) planned by Aya for 2027.

26.3Mining

The PEA study has demonstrated the positive potential of the Boumadine Project. Based on these results, CCE recommends that AYA undertake the following work in preparation for the next phase of the project:

•Conduct further geotechnical investigations to validate the data used for pit slope angles, stope design parameters, and ground support parameters.

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•Locate, survey and model the historical workings (the 1986 to 1992 stoping on four levels and earlier development):

•Conduct a geotechnical risk assessment to define the level of effort required, including void detection surveys where warranted

•Incorporate the workings into the pit and UG designs and the production schedule

•Conduct a hydrogeological investigation to develop a numerical groundwater model to estimate inflows, size the dewatering system and assess pit slope depressurization requirements.

•Confirm the backfill design:

•Determine the strength parameters required for the CRF and develop an optimized mix design that meets the specified mechanical performance and stability criteria

•Build a year-by-year balance of RF and CRF demand against development and OP waste, assessing backhauling and cement logistics costs and the use of surface backfill raises

•Perform mining method trade-off studies, evaluating the following mining methods against modified Avoca:

•Standard Avoca Method

•Longitudinal longhole stoping with paste fill (top-down sequence and bottom-up sequence), including paste plant and tailings requirements

•Validate development advance rates, stope cycle times, and equipment availability and utilization against Zgounder actuals and industry benchmarks and refine crew sizes accordingly.

•Refine estimates of unplanned dilution and mining recovery for the UG mine based on the selected mining method, site-specific geotechnical conditions and backfill properties.

•Complete an OP dilution and mining recovery study:

•Define grade control drilling and blast movement monitoring practices

•Evaluate options to improve selectivity, such as bench height and equipment size, depending on the outcome of the geotechnical study

•Evaluate the UG level spacing (20 m against 25 m and 30 m), considering development requirements, potential rib pillars, dilution and stope stability.

•Conduct a haulage simulation to confirm truck cycle times, fleet size and ramp capacity over the LOM:

•Include all ramp traffic, ramp designs, and shift and blasting constraints

•Confirm the design haul truck and the corresponding haul road widths

•Identify alternatives to improve production, such as truck bypasses, mineralized material and waste passes for mineralized material, haulage levels or shafts

•Define haulage constraints per pit and portal, and refine haulage costs based on the outcome

•Conduct an OP to UG interface trade-off study, evaluating final pit limit against the selected UG mining method, crown pillar location and the incremental stripping ratio, using the finalized parameters from the work described above.

•Conduct a battery-electric versus diesel trade-off study for the UG fleet, considering ventilation demand, electrical consumption, and capital and operating costs.

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•Conduct geochemical testing of the waste rock, including acid-base accounting (ABA), net acid generation (NAG) and kinetic testing.

26.4Processing

The following testwork is recommended to progress the design to the next phase:

•Sample Requirement:

•Spatial composites to represent different physical locations throughout the deposit

•Samples from different zones and depths for variability testing.

•Sufficient sample sizes to conduct downstream testwork.

•Number of representative samples to be determined by AGS.

•Sufficient sample size to generate representative concentrate samples for off-taker evaluation.

•Comminution Testwork:

•Additional tests to determine:

•Crushing work index (CWi)

•SMC parameters: abrasivity (Ai) and impact breakage (A x b)

•BWi (confirmatory)

•Variability of these parameters throughout the deposit and determine correlations to lithology.

•Flotation regrind tests to determine specific energy; engage with vendors for criteria relating to equipment sizing and specifications.

•Ore Characterization Testwork:

•Assess material variability by assaying samples from different zones without blending according to mine plan and to create design ranges.

•Particle size distribution of ROM material and of final concentrates in support of dewatering testwork.

•Rheology testwork to determine apparent viscosities at varying shear rates and varying slurry % solids in support of slurry pipe and pump design.

•Physical characteristics impacting material handling, such as feed moisture content, specific gravity, bulk density, and stacking and clumping / compaction properties.

•Flotation Testwork:

•Variability testing – open circuit and locked cycle to determine the impacts of feed variance and process upsets on grades and recoveries. Develop relations between open circuit and locked cycle results

•Produce enough of each concentrate to support dewatering testwork and further pyrite concentrate treatment testwork.

•Dewatering Testwork:

•Particle settling testwork to determine the flux rate for thickener sizing.

•Flocculant screening tests.

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•Filtration testwork to determine:

•Filter type

•Optimum feed pulp density

•Achievable cake moisture levels

•Transport moisture limit targets

•Filter cycle times

•If any filter aids are required.

•Pyrite Concentrate Treatment Testwork:

•Continue development of pyrite oxidation option.

Assess variability of pyrite concentrate via spatial composites and variability testing.

•Water Analysis:

•Surface water quality analysis.

•Ground water quality analysis.

26.5Tailings Storage Facility

The following recommendations were made that are applicable to the next stage of the study, which have begun since the last PEA update and are currently in progress:

•As per Requirement 5.1, a Multiple Accounts Analysis (MAA) should be conducted to assess alternative tailings technologies in the design of the TSF. The selection of an alternative technology to that described in the PEA may alter the cost estimates for the TSF, as well as for that of the tailings processing and TSF conveyance system.

•The appropriate site location selected for the TSF should be confirmed during the MAA, incorporating the input from experts in the relevant fields, including environmental, social, process, etc, to assess the consequences of hazards posed by the TSF. The selection of an alternative site location to that described in the PEA may alter the cost estimates for the TSF, as well as that of the tailings conveyance system.

The following recommendations are made that are applicable to the next stage of the study, following the MAA:

•The following design aspects of TSF, documented by GCIM, may be reviewed in the next stage of the study and may impact the TSF CAPEX estimate:

•Investigate the suitability of a single 2mm HDPE liner, incorporating the consideration of the permeability of the material upon which it is installed, to mitigate seepage in the case of a liner leak,

•Investigate the suitability of the inclusion of the geotextile beneath the liner. This entails considering both the use as a liner protection measure as well as the resulting seepage in the case of a liner leak,

•Significant floods are generated by the catchments upstream of the TSF. As stated by GCIM, the design of the storm diversion channels is required, adequately sized to cater for the design floods, and

•Sizing of the return water and emergency basins based on a monthly TSF pool water balance.

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26.6Environmental and Social

As part of the ESIA, potential impacts associated with the Project will be identified and assessed in a systematic approach. The impact assessment will inform a comprehensive Environmental and Social Management and Monitoring Plan to be developed for the Project. It will include:

•Mitigation measures for construction, operations and decommissioning

•Framework management plans covering biodiversity, erosion and sediments, waste, air and noise, water, community health and safety, community development and investment, supply chain, cultural heritage, rehabilitation and closure

•Stakeholder engagement plan including grievance management

•Monitoring plans for Project performance and the receiving environment (water, air, noise and biodiversity) including social aspects

Water management planning will continue to include the following elements:

•Characterization of the groundwater systems on site

•Environmental testwork to understand the geochemical composition and acid generating potential of the host rock and mineral residues

•Development of a site-wide climatic water balance to understand the water circuit and confirm water use, storage and discharge volume requirements.

•Establishment of an engineered stormwater runoff system catering for the diversion of clean water runoff around Project sites, separation of clean and potentially dirty water runoff on site, storage of potentially dirty water runoff in appropriately designed and engineered facilities, and collection and re-use of potentially dirty water runoff.

•Flood protection analysis and measures, if needed.

•Prioritizing water reuse and recycling.

•Water treatment for potable water supply, raw water supply, and discharge to the environment (if discharge is required).

26.7Project Next Phases

To progress the project, it is recommended to conduct pre-feasibility level trade offs, design, and engineering to improve the accuracy of cost estimation. This shall include optimizing the mine plan, updating the process design using results from additional metallurgical test data, and conduct engineering of other disciplines such as mechanical, civil, structural and electrical engineering producing MTOs. The information from this phase of project will then feed into a Feasibility Study.

The advancement to a Feasibility Study will be dependent on project economics from pre-feasibility level designs and progress of drilling to advance classification of resources.

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27.0REFERENCES

•Abia, E.H., Nachit, H., Ibhi, A. and Baroudi, Z. 1999. Les Minéralisations Filoniennes à Pb, Zn et Cu de la Boutonnière de l'Ougnat. Relations avec les Déformations et Essai de Calage Chronologique. Chronique Recherche Minière 536, p. 83-95.

•Abia, E.H., Nachit, H., Christian Marignac, C., Abderrahmane, I. and Saadi, S.A. 2003. The Polymetallic Au–Ag-Bearing Veins of Bou Madine (Jbel Ougnat, eastern Anti-Atlas, Morocco): Tectonic Control and Evolution of a Neoproterozoic Epithermal Deposit. Journal of African Earth Sciences 36, 251-271.

•Ait Saadi. 1992. Contribution à l’étude de l’environnement paléovolcanique du Protérozoique Supérieur (PIII) et du mode de genèse des concentrations polymétalliques (Zn, Pb, Cu, Ag, Au) de Bou Madine (Ougnat, Anti-Atlas oriental, Maroc). Unpublished Thesis, University of INPL, Nancy, p. 218.

•Alansari A. and Sagon JP. 1997. Le gisement de Tiouit (Jebel Saghro, Anti-Atlas, Maroc): un système mésothermal polyphasé à sulfures-or et hématite-or dans une granodiorite potassique d’âge Protérozoîque. Chron Recherche Minière 527, 3-25.

•Alfieris, D. Voudouris, P. and Spry, P.G. 2013. Shallow submarine epithermal Pb-Zn-Cu-Au-Ag-Te mineralization on western Milos Island, Aegean Volcanic Arc, Greece: Mineralogical, geological and geochemical constraints. Ore Geology Reviews 53, 159-180.

•Aya Gold & Silver 2025. NI 43-101 Technical Report and Updated Mineral Resource Estimate of the Boumadine Polymetallic Project, Kingdom of Morocco. Prepared for Aya Gold & Silver Inc. by Lalonde, D, and Pérez, P. Effective Date: February 24, 2025. Issue Date : March 31, 2025. 193 p.

•Aya Gold & Silver 2025b. NI 43-101 Technical Report. Preliminary Economic Assessment for the Boumadine Polymetallic Project, Kingdom of Morrocco. Prepared for Aya Gold & Silver Inc. Prepared by Lalonde, D, Pérez, P, Beaudoin, R, Humphrey, Nayak, P., Venter, R., Shaikh, Z., Berson, B., Pheiffer, A., Papageorgiou, G., Puritch, E., Yassa, A., Brown, F., Barry, J., Stone, W. and Palleske, C R. Effective Date: November 4, 2025. Issue Date: December 18, 2025. 343 p.

•Aya Gold & Silver 2025b. NI 43-101 Technical Report. Updated Mineral Resource and ineral Reserve Assessment for the Zgounder Silver Mine Operation, Kingdom of Morrocco. Prepared for Aya Gold & Silver Inc. Prepared by Lalonde, D, Pérez, P, Beaudoin, R, Humphrey, S, vilder, D, Preetham, N, Venter, R. Issue Date: December 16, 2025.. 337 p.

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•BGRIMM Technology Group. 2025. Research Report on the Small-Scale Test Project on Roasting and Cyanidation of Boumadine Pyrite Concentrate. Prepared for Boumadine Global Mining S.A. Dated October 22, 2025. 67 pages.

•Boily, M. 2014. The Boumadine Polymetallic (Au, Ag, Zn, Pb, Cu) Deposit Errachidia Province, Kingdom of Morocco. Prepared for Maya Gold and Silver Inc. by Michel Boily, Ph.D., P.Geo. of Geon Ltee, dated April 2, 2014. 217 pages.

•Bouabdellah, M., Chekroun, F., Alansari, A. and Margoum, D. 2016. The Granitoid-Related Tiouit Gold Deposit, Saghro Inlier, Eastern Anti-Atlas (Morocco): Neoproterozoic Mineralization by a polyphase Late-Magmatic to Hydrothermal System. In M. Bouabdellah and J.F. Slack (eds), Mineral Deposits of North Africa.

•Bouabdellah, M. and Levresse, G. 2016. The Boumâadine Polymetallic Ore Deposit, Eastern Anti-Atlas, Morocco: Evolution from massive Fe-As-Sn to epithermal Au-Ag-Zn±Cu mineralization in a Neoproterozoic

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resurgent Caldera environment. In Mineral Deposits of North Africa: Cham, Switzerland, Springer International Publishing, 133–142.

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•BRPM. 1998. Gisement Polymétallique de Boumadine. Géologie et États des Réserves. BRPM, 30 p.

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•Cheilletz, A., Levresse, G., Gasquet, D., Azizi Samir, M.R., Zyadi, R. and Archibald, D. 2002. The Imiter epithermal silver deposit: a giant Neoproterozoic epithermal mineralization in the Anti-Atlas, Morocco. Mineralium Deposita 37, 772–781.

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•Epoch, Pre-Economic Assessment of the Boumadine Tailings Storage Facilities, Report No.: 255-100-REP-0-02-10-2024, Submitted to Aya Silver & Gold Inc. on December 9, 2024.

•Epoch, Updated to the Pre-Economic Assessment of the Boumadine Tailings Storage Facilities, Report No.: 255-100-REP-0-02-10-2024, Submitted to Aya Silver & Gold Inc. on November 20, 2025.

•Freton, R. 1988. Contribution à l’Étude Métallogénique ou District de Bou-Madine Anti-Atlas, Maroc, Environnement Tectonique et Concentrations Épithermales B.P.G.C à Ag et Au. Thèse de Doctorat, Institut National Polytechnique de Lorraine de Géologie Appliquée et de Prospection Minière, 63 pages.

•Gaboury, D. (2022a). Gisement polymétallique Boumadine (Au-Ag-Zn-Pb), Maroc : Contexte géologique et implications pour l’exploration. Rapport interne soumis à Aya Gold and Silver, 12 May 2022, 40 p.

•Gaboury, D. (2022b). Gisement polymétallique Boumadine (Au-Ag-Zn-Pb), Maroc : Revue de sondages sélectionnés de 2022 et bordures Sud et Ouest de la caldeira. Rapport interne soumis à Aya Gold and Silver, 15 December 2022, 52 p

•Gaboury, D. (2023a). Gisement polymétallique Boumadine (Au-Ag-Zn-Pb), Maroc : facteurs d’enrichissement en Zn et Au des minéralisations à sulfures. Report prepared for Aya Gold & Silver Inc., final version dated 5 August 2023, 50 p.

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•Gaboury, D. (2023b). Modèle régional de minéralisation à proximité du gisement polymétallique Boumadine (Au-Ag-Zn-Pb), Maroc. Report prepared for Aya Gold & Silver Inc., first version dated 22 November 2023, 40 p.

•Gaboury, D. (2023c). Transition latérale des styles de minéralisations polymétalliques, gisement Boumadine (Au-Ag-Zn-Pb), Maroc. Report prepared for Aya Gold & Silver Inc., second version dated 29 December 2023, 31 p.

•Gaboury, D. (2025a). Rapport sur les forages aux extrêmes Nord, Sud et Ouest, et en périphérie du BOU-DD22-036, gisement Boumadine (Au-Ag-Zn-Pb), Maroc. Report prepared for Aya Gold & Silver Inc., final version dated 5 June 2025, 28 p.

•Gaboury, D. (2025b). Étude des forages : BOU-DD25-600 (exploration Tanda), BOU-DD25-564 (forage profond zone centre), BOU-DD24-461, 463 (zones anomaliques en or). Report prepared for Aya Gold & Silver Inc., preliminary version dated 30 June 2025, 24 p.

•Gaboury, D. (2025c). Mise en contexte géologique et potentiel des permis régionaux, projet Boumadine, Maroc. Report prepared for Aya Gold & Silver Inc., final version dated 14 August 2025, 80 p.GCIM, Boumadine Project, TSF Preliminary Economic Assessment - Rapport du PEA du TSF Flottation Option 18.6 MT. September 17, 2025.

•GCIM, Boumadine Project, TSF Preliminary Economic Assessment - Rapport du PEA du TSF Flottation Option 18.6 MT. September 17, 2025

•GCIM, Boumadine Project, Alimentation en eau du Projet Boumadine - Evaluation Économique Préliminaire - Rapport Alimentation en eau. October 31, 2024

•Hannington, M.D., De Ronde, C.E.J. and Petersen, S. 2005. Seafloor Tectonics and Submarine Hydrothermal Systems. Economic Geology 100th Anniversary Volume, 111-142.

•Hefferan K, Soulaimani A, Samson SD et al. 2014. A reconsideration of Pan African orogenic cycle in the Anti-Atlas Mountains, Morocco. Journal of African Earth Sciences 98, 34–46. Doi:10.1016/j. jafrearsci.2014.03.007 Paile Y., Paris XI , Orsay, France, (1983) 290.

•Leistel, J.M. and Qadrouci, A. 1991. Le gisement argentifère d’Imiter (Protérozoîque supérieur de l’Anti-Atlas, Maroc). Contrôles des minéralisations, hypothèse génétique et perspectives pour l’exploration. Chron Rech Min 502, 5–22.

•Lelièvre, J. 2011. Essais minéralurgiques sur les résidus miniers du site Boumadine, Maroc – URSTM Rapport PU-2011-02-586.

•Levresse, G. 2001. Contribution à l’établissement d’un modèle génétique des gisements d’Imiter (Ag–Hg), Bou Madine (Pb –Zn –Cu –Ag–Au) et Bou Azzer (Co –Ni –As–Au –Ag) dans l’Anti-Atlas marocain. Unpubl. Ph.D. thesis, INPL, Nancy, France.

•Levresse, G., Cheilletz, A., Gasquet, D., Reisberg, L., Deloule, E., Marty, B. and Kyser, K. 2004. Osmium, sulphur, and helium isotopic results from the giant Neoproterozoic epithermal Imiter silver deposit, Morocco: evidence for a mantle source. Chemical Geology 207, 59-79.

•Maya – Boumadine – 500 TPD Polymetallic Ore Flotation Processing Flow – PDF File Shandong Xinhai Mining Machinery, No date.

•Mouttaqi, A., Rjimati, E.C. ., Maacha, L., Michard, A., Soulaimani, A. and Ibouh, A. 2001. LES PRINCIPALES MINES DU MAROC/Main Mines of Morocco, NOTES ET MÉMOIRES DU SERVICE GÉOLOGIQUE N° 564 Volume 9.

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•Ouguir, H., Macaudière J., Dagallier G., Qadrouci A. and Leistel J.M. 1994. Cadre structural du gîte Ag-Hg d’Imiter (Anti-Atlas, Maroc); Implication métallogénique. Bull. Soc. Géol. Fr. 165, 233–248.

•P&E Mining Consultants Inc. 2024. NI 43-101 Technical Report and updated Mineral Resource Estimate of the Boumadine Polymetallic Project, Kingdom of Morocco. Prepared for Aya Gold & Silver Inc. by Stone, W., Brown, F., Barry, J., Yassa, A., Feasby, G. and Puritch, E. Effective Date: May 08, 2024. Issue Date : May 31, 2024. 214 p.

•Paile, Y. 1983. Etude des séries volcaniques du Précambrien III de l'Ougnat (Anti-Atlas oriental, Maroc) et de ses minéralisations plombo-zincifères complexes associées (Gîte de Bou Madine), Thèse 3ème cycle, Université Paris, Orsay. 290 pages.

•Popov, A. and Belkasmi, A. 1983. Gisement de cuivre d’Oumjrane-Boukerzia. Rapport interne BRPM, 7 p. in Mouttaqi et al., 2001.

•Rapport sur les essais en laboratoire sur les échantillons du résidu de la mine Boumadine au Maroc. Nichromet Extraction Inc. 31 mai 2013.

•Read, J., & Stacey, P. (2009). Guidelines for open pit slope design. 496p. CRC Press/Balkema: Leiden.

•RockEng, Boumadine Project PEA Geomechanics. Report #24017-102. Submitted to Aya Silver & Gold Inc. on 30 January 2025.

•SCX Technology, Étude logistique du transport en vrac de minerai , October 2025

•SGS. 2018a. Memo Report, June 20, 2018.

•SGS. 2018b. An Investigation into the recovery of gold and silver from Boumâadine. Prepared for Maya Gold & Silver Inc - Project 16525-001A and 16525-02 – Final Report – December 18, 2018.

•SGS. 2022. An investigation into Metallurgical Testing of a Main Composite from the Boumadine Deposit. Prepared for Aya Gold & Silver Inc – Project 18098-01 – Final Report – Revision 1. Dated July 14, 2022. 233 pages.

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•Sillitoe, R.H. 2010. Porphyry copper systems. Economic Geology 105, 3-41.

•Soulaimani A, Michard A, Ouanaimi H et al. 2014. Late Ediacaran–Cambrian structures and their reactivation during the Variscan and Alpine cycles in the Anti-Atlas (Morocco). Journal of African Earth Sciences 98, 94–112. Doi :10.1016/j.jafrearsci.2014.04.025.

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28.0CERTIFICATES

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Appendix AQAQC CRM PLOTS

CRM plots with certified values include only, includes four acid digestion. In some cases CRM’s with extreme values relating to mis-labelling or sample swaps have been removed from the images below.

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Appendix BSTATISTICAL VALIDATION - UNCAPPED VS CAPPED COMPOSITES

Uncapped composites vs capped composites

Domain			Element			Uncapped comps mean			Uncapped comps COV			Capped comps mean			Capped comps COV			St difference (mean)			% Difference (mean)			Straight difference (COV)
1003			Ag (g/t)			44.31			2.25			44.31			2.25			0			0			0
1003			Au (g/t)			0.56			1.25			0.56			1.25			0			0			0
1003			Pb (%)			0.42			1.69			0.42			1.69			0			0			0
1003			Zn (%)			1.87			2.28			1.87			2.28			0			0			0
1003			Cu (%)			0.09			1.5			0.09			1.49			0			0			-0.01
1003			Fe (%)			17.31			1.74			17.31			1.74			0			0			0
1004			Ag (g/t)			65.42			0.93			65.42			0.93			0			0			0
1004			Au (g/t)			0.35			0.92			0.35			0.92			0			0			0
1004			Pb (%)			1.79			1.27			1.79			1.27			0			0			0
1004			Zn (%)			2.57			0.71			2.57			0.71			0			0			0
1004			Cu (%)			0.06			1.4			0.06			1.4			0			0			0
1004			Fe (%)			14.52			0.87			13.01			0.76			-1.51			-0.1			-0.1
1005			Ag (g/t)			73.82			0.92			73.82			0.92			0			0			0
1005			Au (g/t)			0.47			0.53			0.47			0.53			0			0			0
1005			Pb (%)			0.27			0.37			0.27			0.37			0			0			0
1005			Zn (%)			1.05			0.37			1.05			0.37			0			0			0
1005			Cu (%)			0.03			0.9			0.03			0.9			0			0			0
1005			Fe (%)			10.88			0.77			10.88			0.77			0			0			0
1006			Ag (g/t)			30.75			0.65			30.75			0.65			0			0			0
1006			Au (g/t)			0.38			0.8			0.38			0.8			0			0			0
1006			Pb (%)			0.45			0.74			0.45			0.74			0			0			0
1006			Zn (%)			1.45			0.75			1.45			0.75			0			0			0
1006			Cu (%)			0.02			0.71			0.02			0.71			0			0			0
1006			Fe (%)			11.99			0.97			11.99			0.97			0			0			0
1007			Ag (g/t)			200.32			1.61			177.74			1.57			-22.58			-0.11			-0.05
1007			Au (g/t)			0.64			0.58			0.64			0.58			0			0			0
1007			Pb (%)			0.53			0.94			0.53			0.94			0			0			0
1007			Zn (%)			2.51			1.37			2.51			1.37			0			0			0
1007			Cu (%)			0.23			1.41			0.23			1.41			0			0			0
1007			Fe (%)			14.65			0.63			13.61			0.55			-1.04			-0.07			-0.08
1011			Ag (g/t)			57.16			0.72			57.16			0.72			0			0			0
1011			Au (g/t)			1.59			0.95			1.59			0.95			0			0			0
1011			Pb (%)			0.47			0.69			0.47			0.69			0			0			0
1011			Zn (%)			0.92			0.73			0.92			0.73			0			0			0
1011			Cu (%)			0.03			0.89			0.03			0.89			0			0			0
1011			Fe (%)			35.59			0.87			35.59			0.87			0			0			0
1012			Ag (g/t)			129.23			0.38			129.23			0.38			0			0			0
1012			Au (g/t)			1			0.86			1			0.86			0			0			0
1012			Pb (%)			1.24			0.86			1.24			0.86			0			0			0
1012			Zn (%)			2.26			0.93			2.26			0.93			0			0			0
1012			Cu (%)			0.31			0.63			0.31			0.63			0			0			0
1012			Fe (%)			22.78			0.46			22.78			0.46			0			0			0
1013			Ag (g/t)			29.75			0.7			29.75			0.7			0			0			0

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        427

---

1013			Au (g/t)			0.12			0.91			0.12			0.91			0			0			0
1013			Pb (%)			0.49			1.34			0.49			1.34			0			0			0
1013			Zn (%)			1.07			1.39			1.07			1.39			0			0			0
1013			Cu (%)			0.03			0.38			0.03			0.38			0			0			0
1013			Fe (%)			17.39			0.76			17.39			0.76			0			0			0
1030			Ag (g/t)			130.13			1.25			130.13			1.25			0			0			0
1030			Au (g/t)			2.29			1.92			2.29			1.92			0			0			0
1030			Pb (%)			1.64			1.2			1.64			1.2			0			0			0
1030			Zn (%)			7.65			1.37			7.65			1.37			0			0			0
1030			Cu (%)			0.21			1.85			0.21			1.85			0			0			0
1030			Fe (%)			21.76			1.09			21.76			1.09			0			0			0
1031			Ag (g/t)			99.97			1.34			96.85			1.32			-3.13			-0.03			-0.02
1031			Au (g/t)			1.31			3.26			1.31			3.26			0			0			0
1031			Pb (%)			0.88			1.36			0.88			1.36			0			0			0
1031			Zn (%)			2.43			2.11			2.43			2.11			0			0			0
1031			Cu (%)			0.06			2.25			0.06			2.25			0			0.01			0
1031			Fe (%)			19.63			1.85			19.63			1.85			0			0			0
1033			Ag (g/t)			20.62			0.83			20.62			0.83			0			0			0
1033			Au (g/t)			0.99			0.96			0.99			0.96			0			0			0
1033			Pb (%)			0.24			0.71			0.24			0.71			0			0			0
1033			Zn (%)			0.36			1.06			0.36			1.06			0			0			0
1033			Cu (%)			0.03			1.57			0.03			1.54			0			0.02			-0.03
1033			Fe (%)			12.42			0.7			12.42			0.7			0			0			0
1040			Ag (g/t)			85.8			1.05			85.8			1.05			0			0			0
1040			Au (g/t)			2.58			1.34			2.45			1.34			-0.12			-0.05			0
1040			Pb (%)			2.14			1.83			2.14			1.83			0			0			0
1040			Zn (%)			3.84			2.07			3.84			2.07			0			0			0
1040			Cu (%)			0.06			1.07			0.06			1.06			0			0.01			0
1040			Fe (%)			24.28			0.81			24.28			0.81			0			0			0
1041			Ag (g/t)			148.18			1.49			148.18			1.49			0			0			0
1041			Au (g/t)			1.74			2.08			1.67			2.12			-0.08			-0.04			0.05
1041			Pb (%)			2.16			1.49			2.16			1.49			0			0			0
1041			Zn (%)			3.57			1.43			3.49			1.43			-0.08			-0.02			0
1041			Cu (%)			0.04			1.26			0.04			1.26			0			0			0
1041			Fe (%)			19.96			0.96			19.96			0.96			0			0			0
1050			Ag (g/t)			46.76			1.27			46.76			1.27			0			0			0
1050			Au (g/t)			2.65			1.5			2.65			1.5			0			0			0
1050			Pb (%)			0.93			2.06			0.93			2.06			0			0			0
1050			Zn (%)			2.91			2.03			2.75			1.97			-0.17			-0.06			-0.06
1050			Cu (%)			0.35			1.48			0.35			1.48			0			0			0
1050			Fe (%)			22.58			1.29			22.58			1.29			0			0			0
1051			Ag (g/t)			27.98			0.81			27.98			0.81			0			0			0
1051			Au (g/t)			0.3			0.46			0.3			0.46			0			0			0
1051			Pb (%)			0.11			0.63			0.11			0.63			0			0			0
1051			Zn (%)			0.04			0.91			0.04			0.91			0			0			0
1051			Cu (%)			0.01			0.84			0.01			0.73			0			0.06			-0.11
1051			Fe (%)			3.67			0.1			3.67			0.1			0			0			0
1052			Ag (g/t)			48.42			1.37			48.42			1.37			0			0			0
1052			Au (g/t)			0.78			1.32			0.78			1.32			0			0			0
1052			Pb (%)			0.82			1.23			0.82			1.23			0			0			0
1052			Zn (%)			0.87			1.27			0.87			1.27			0			0			0

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        428

---

1052			Cu (%)			0.19			2.84			0.11			1.39			-0.08			-0.41			-1.45
1052			Fe (%)			17.67			0.98			17.67			0.98			0			0			0
1053			Ag (g/t)			39.73			0.57			39.73			0.57			0			0			0
1053			Au (g/t)			0.84			0.3			0.84			0.3			0			0			0
1053			Pb (%)			0.68			1.04			0.68			1.04			0			0			0
1053			Zn (%)			1.92			1.3			1.92			1.3			0			0			0
1053			Cu (%)			0.07			0.51			0.07			0.51			0			0			0
1053			Fe (%)			13.65			0.3			13.65			0.3			0			0			0
1054			Ag (g/t)			22.97			1.34			22.97			1.34			0			0			0
1054			Au (g/t)			0.05			1.19			0.05			1.19			0			0			0
1054			Pb (%)			0.97			1.7			0.97			1.7			0			0			0
1054			Zn (%)			1.98			2.14			1.98			2.14			0			0			0
1054			Cu (%)			0.25			0.86			0.25			0.86			0			0			0
1054			Fe (%)			9.14			1.48			9.14			1.48			0			0			0
1055			Ag (g/t)			296.3			0.68			296.3			0.68			0			0			0
1055			Au (g/t)			0.72			0.7			0.72			0.7			0			0			0
1055			Pb (%)			4.74			0.7			4.74			0.7			0			0			0
1055			Zn (%)			4.53			0.85			4.53			0.85			0			0			0
1055			Cu (%)			0.07			0.92			0.07			0.92			0			0			0
1055			Fe (%)			20.67			0.71			20.67			0.71			0			0			0
1056			Ag (g/t)			49.3			0.64			49.3			0.64			0			0			0
1056			Au (g/t)			0.71			0.98			0.71			0.98			0			0			0
1056			Pb (%)			0.69			1.57			0.69			1.57			0			0			0
1056			Zn (%)			0.65			1.12			0.65			1.12			0			0			0
1056			Cu (%)			0.46			0.61			0.46			0.61			0			0			0
1056			Fe (%)			40.23			0.91			37.39			0.81			-2.84			-0.07			-0.1
1061			Ag (g/t)			122.19			0.6			122.19			0.6			0			0			0
1061			Au (g/t)			0.62			0.78			0.62			0.78			0			0			0
1061			Pb (%)			2.82			0.59			2.82			0.59			0			0			0
1061			Zn (%)			6.75			0.85			6.75			0.85			0			0			0
1061			Cu (%)			0.03			0.11			0.03			0.03			0			0.14			-0.08
1061			Fe (%)			28.03			0.52			28.03			0.52			0			0			0
1062			Ag (g/t)			22.45			1.35			22.45			1.35			0			0			0
1062			Au (g/t)			0.06			0.85			0.06			0.85			0			0			0
1062			Pb (%)			5.21			1.41			5.21			1.41			0			0			0
1062			Zn (%)			10.88			1.41			9.91			1.41			-0.97			-0.09			0
1062			Cu (%)			0.22			0.66			0.22			0.66			0			0.01			0
1062			Fe (%)			17.55			0.74			17.55			0.74			0			0			0
1063			Ag (g/t)			3.08			0.77			3.08			0.77			0			0			0
1063			Au (g/t)			7.94			0.48			7.94			0.48			0			0			0
1063			Pb (%)			0.02			0.64			0.02			0.64			0			0			0
1063			Zn (%)			0.03			0.86			0.03			0.86			0			0			0
1063			Cu (%)			0.02			0.54			0.02			0.54			0			0			0
1063			Fe (%)			34.44			0.83			34.44			0.83			0			0			0
1090			Ag (g/t)			258.26			1.94			258.26			1.94			0			0			0
1090			Au (g/t)			3.77			1.31			3.77			1.31			0			0			0
1090			Pb (%)			2.87			2.39			2.87			2.39			0			0			0
1090			Zn (%)			9.41			1.89			9.35			1.88			-0.06			-0.01			-0.01
1090			Cu (%)			0.24			1.93			0.24			1.93			0			0			-0.01
1090			Fe (%)			44.69			0.99			43.29			0.89			-1.4			-0.03			-0.1
1091			Ag (g/t)			432.92			1.61			394.35			1.27			-38.57			-0.09			-0.34

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        429

---

1091			Au (g/t)			1.68			1.17			1.65			1.11			-0.04			-0.02			-0.06
1091			Pb (%)			3.07			1.27			2.74			1.04			-0.33			-0.11			-0.23
1091			Zn (%)			4.68			1.04			4.43			0.92			-0.25			-0.05			-0.12
1091			Cu (%)			0.14			2.22			0.14			2.13			0			-0.03			-0.09
1091			Fe (%)			32.27			0.77			32.27			0.77			0			0			0
1092			Ag (g/t)			91.96			1.89			91.96			1.89			0			0			0
1092			Au (g/t)			4.19			1.9			4.19			1.9			0			0			0
1092			Pb (%)			1.25			1.14			1.25			1.14			0			0			0
1092			Zn (%)			3.2			1.75			3.2			1.75			0			0			0
1092			Cu (%)			0.27			1.67			0.27			1.66			0			0			0
1092			Fe (%)			43.15			1.27			43.15			1.27			0			0			0
1093			Ag (g/t)			329.44			2.49			329.44			2.49			0			0			0
1093			Au (g/t)			2.51			1.53			2.51			1.53			0			0			0
1093			Pb (%)			1.04			1.3			1.04			1.3			0			0			0
1093			Zn (%)			3.26			1.66			3.26			1.66			0			0			0
1093			Cu (%)			0.22			1.55			0.22			1.55			0			0			0
1093			Fe (%)			29.68			0.95			29.68			0.95			0			0			0
1094			Ag (g/t)			105.6			1.04			105.6			1.04			0			0			0
1094			Au (g/t)			6.21			1.2			6.21			1.2			0			0			0
1094			Pb (%)			1.19			0.77			1.19			0.77			0			0			0
1094			Zn (%)			1.51			1.11			1.51			1.11			0			0			0
1094			Cu (%)			0.2			1.24			0.2			1.23			0			0			0
1094			Fe (%)			49.41			0.71			49.41			0.71			0			0			0
1201			Ag (g/t)			1095.83			1.65			1058.78			1.64			-37.05			-0.03			-0.01
1201			Au (g/t)			5.78			1.17			5.78			1.17			0			0			0
1201			Pb (%)			10.26			1.72			10.21			1.73			-0.05			-0.01			0.01
1201			Zn (%)			12.92			1.01			12.87			1.01			-0.05			0			0
1201			Cu (%)			0.22			0.93			0.22			0.93			0			0			0
1201			Fe (%)			78.75			0.88			78.75			0.88			0			0			0
1202			Ag (g/t)			110.65			2.3			110.65			2.3			0			0			0
1202			Au (g/t)			1.96			2.99			1.63			2.84			-0.33			-0.17			-0.15
1202			Pb (%)			1.17			2.25			1.17			2.25			0			0			0
1202			Zn (%)			2.74			1.89			2.74			1.89			0			0			0
1202			Cu (%)			0.1			2.71			0.1			2.7			0			0			-0.01
1202			Fe (%)			19.41			1.52			14.19			1.09			-5.21			-0.27			-0.43
1203			Ag (g/t)			71.92			2.04			71.92			2.04			0			0			0
1203			Au (g/t)			1.48			2.78			1.48			2.78			0			0			0
1203			Pb (%)			1.4			1.96			1.4			1.96			0			0			0
1203			Zn (%)			3.14			1.93			3.14			1.93			0			0			0
1203			Cu (%)			0.05			1.59			0.05			1.58			0			0			0
1203			Fe (%)			18.33			1.61			18.33			1.61			0			0			0
1208			Ag (g/t)			21.12			0.65			21.12			0.65			0			0			0
1208			Au (g/t)			0.58			0.32			0.58			0.32			0			0			0
1208			Pb (%)			0.16			0.71			0.16			0.71			0			0			0
1208			Zn (%)			0.16			0.97			0.16			0.97			0			0			0
1208			Cu (%)			0.04			0.89			0.04			0.87			0			0.01			-0.02
1208			Fe (%)			13.18			0.58			13.18			0.58			0			0			0
1209			Ag (g/t)			20.16			0.83			20.16			0.83			0			0			0
1209			Au (g/t)			0.57			1.74			0.57			1.74			0			0			0
1209			Pb (%)			0.36			0.7			0.36			0.7			0			0			0
1209			Zn (%)			1.34			0.82			1.34			0.82			0			0			0

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        430

---

1209			Cu (%)			0.18			1.75			0.18			1.75			0			0			0
1209			Fe (%)			7.77			0.58			7.77			0.58			0			0			0
1304			Ag (g/t)			29.04			0.4			29.04			0.4			0			0			0
1304			Au (g/t)			1.12			0.79			1.12			0.79			0			0			0
1304			Pb (%)			1.41			0.99			1.41			0.99			0			0			0
1304			Zn (%)			1.28			0.84			1.28			0.84			0			0			0
1304			Cu (%)			0.07			0.4			0.07			0.37			0			0.02			-0.02
1304			Fe (%)			24.37			0.63			24.37			0.63			0			0			0
1305			Ag (g/t)			54.67			1.31			54.67			1.31			0			0			0
1305			Au (g/t)			2.3			1.54			2.3			1.54			0			0			0
1305			Pb (%)			0.58			1.46			0.58			1.46			0			0			0
1305			Zn (%)			1.03			1.38			1.03			1.38			0			0			0
1305			Cu (%)			0.16			1.05			0.16			1.05			0			0			0
1305			Fe (%)			31.91			0.82			31.06			0.75			-0.85			-0.03			-0.07
1306			Ag (g/t)			144.14			1.15			144.14			1.15			0			0			0
1306			Au (g/t)			2.74			1.74			2.74			1.74			0			0			0
1306			Pb (%)			1.37			1.13			1.37			1.13			0			0			0
1306			Zn (%)			2.48			1.32			2.48			1.32			0			0			0
1306			Cu (%)			0.21			1.45			0.21			1.45			0			0			0
1306			Fe (%)			31.39			1.19			31.39			1.19			0			0			0
1310			Ag (g/t)			46.17			0.9			46.17			0.9			0			0			0
1310			Au (g/t)			2.67			1.27			2.67			1.27			0			0			0
1310			Pb (%)			0.99			1.3			0.99			1.3			0			0			0
1310			Zn (%)			2.32			1.67			2.32			1.67			0			0			0
1310			Cu (%)			0.17			1.16			0.17			1.16			0			0			0
1310			Fe (%)			33.37			0.95			33.37			0.95			0			0			0
1311			Ag (g/t)			43.29			1.11			43.29			1.11			0			0			0
1311			Au (g/t)			1.31			2.01			1.31			2.01			0			0			0
1311			Pb (%)			0.12			0.62			0.12			0.62			0			0			0
1311			Zn (%)			0.43			0.76			0.43			0.76			0			0			0
1311			Cu (%)			0.1			0.96			0.1			0.96			0			0			0
1311			Fe (%)			32.96			1.12			27.39			0.91			-5.57			-0.17			-0.21
1312			Ag (g/t)			44.6			1.78			44.6			1.78			0			0			0
1312			Au (g/t)			0.83			1.78			0.83			1.78			0			0			0
1312			Pb (%)			0.58			0.96			0.58			0.96			0			0			0
1312			Zn (%)			3.18			2.43			3.18			2.43			0			0			0
1312			Cu (%)			0.1			1.61			0.1			1.58			0			0.01			-0.03
1312			Fe (%)			15.19			1.13			15.19			1.13			0			0			0
1313			Ag (g/t)			38.1			1.89			38.1			1.89			0			0			0
1313			Au (g/t)			0.21			1.02			0.21			1.02			0			0			0
1313			Pb (%)			2.75			2.07			2.75			2.07			0			0			0
1313			Zn (%)			2.44			0.86			2.44			0.86			0			0			0
1313			Cu (%)			0.04			0.87			0.04			0.89			0			0.04			0.01
1313			Fe (%)			10.56			0.5			10.56			0.5			0			0			0
1314			Ag (g/t)			24.01			0.88			24.01			0.88			0			0			0
1314			Au (g/t)			1.25			0.92			1.25			0.92			0			0			0
1314			Pb (%)			0.32			1.22			0.32			1.22			0			0			0
1314			Zn (%)			0.49			1.55			0.49			1.55			0			0			0
1314			Cu (%)			0.12			0.91			0.12			0.9			0			0			-0.01
1314			Fe (%)			23.93			0.5			23.93			0.5			0			0			0
1331			Ag (g/t)			39.9			0.84			39.9			0.84			0			0			0

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        431

---

1331			Au (g/t)			1.49			1.84			1.49			1.84			0			0			0
1331			Pb (%)			2.13			1.74			2.13			1.74			0			0			0
1331			Zn (%)			2.18			0.93			2.18			0.93			0			0			0
1331			Cu (%)			0.11			1.43			0.11			1.43			0			0.01			-0.01
1331			Fe (%)			21.34			0.84			21.34			0.84			0			0			0
1341			Ag (g/t)			41.69			1.47			41.69			1.47			0			0			0
1341			Au (g/t)			0.75			1.52			0.75			1.52			0			0			0
1341			Pb (%)			2.68			2.45			2.68			2.45			0			0			0
1341			Zn (%)			4.24			1.69			4.24			1.69			0			0			0
1341			Cu (%)			0.12			1.8			0.12			1.79			0			0.01			-0.01
1341			Fe (%)			18.34			0.81			18.34			0.81			0			0			0
1403			Ag (g/t)			36.95			0.81			36.95			0.81			0			0			0
1403			Au (g/t)			0.89			1.45			0.89			1.45			0			0			0
1403			Pb (%)			3.68			1.42			3.68			1.42			0			0			0
1403			Zn (%)			3.36			0.74			3.36			0.74			0			0			0
1403			Cu (%)			0.09			0.65			0.09			0.65			0			0.01			0
1403			Fe (%)			32.88			0.55			32.88			0.55			0			0			0
1404			Ag (g/t)			22.84			1.17			22.84			1.17			0			0			0
1404			Au (g/t)			0.21			1.37			0.21			1.37			0			0			0
1404			Pb (%)			0.39			0.83			0.39			0.83			0			0			0
1404			Zn (%)			1.39			1.27			1.39			1.27			0			0			0
1404			Cu (%)			0.06			2.39			0.07			2.36			0			0.01			-0.02
1404			Fe (%)			11.09			0.58			11.09			0.58			0			0			0
1407			Ag (g/t)			45			1.25			45			1.25			0			0			0
1407			Au (g/t)			2.57			2.62			1.99			2.26			-0.58			-0.22			-0.36
1407			Pb (%)			0.98			1.3			0.98			1.3			0			0			0
1407			Zn (%)			1.76			1.18			1.76			1.18			0			0			0
1407			Cu (%)			0.15			1.55			0.15			1.54			0			0.01			-0.01
1407			Fe (%)			39.15			1.21			39.15			1.21			0			0			0
1501			Ag (g/t)			207.96			1.23			207.96			1.23			0			0			0
1501			Au (g/t)			2.23			1.2			2.14			1.14			-0.09			-0.04			-0.06
1501			Pb (%)			2.64			1.27			2.64			1.27			0			0			0
1501			Zn (%)			5.15			1.32			5.15			1.32			0			0			0
1501			Cu (%)			0.08			1.19			0.08			1.18			0			0.01			-0.01
1501			Fe (%)			37.53			0.89			37.53			0.89			0			0			0
1502			Ag (g/t)			275.26			1.53			275.26			1.53			0			0			0
1502			Au (g/t)			1.2			1.02			1.2			1.02			0			0			0
1502			Pb (%)			2.51			1.26			2.51			1.26			0			0			0
1502			Zn (%)			6.3			1.23			6.19			1.24			-0.12			-0.02			0.01
1502			Cu (%)			0.06			1.49			0.06			1.48			0			0.01			-0.01
1502			Fe (%)			24.74			0.86			24.74			0.86			0			0			0
1503			Ag (g/t)			236.47			1.4			230.2			1.36			-6.27			-0.03			-0.04
1503			Au (g/t)			1.2			1.04			1.2			1.04			0			0			0
1503			Pb (%)			2.48			1.52			2.38			1.44			-0.1			-0.04			-0.07
1503			Zn (%)			4.92			1.32			4.92			1.32			0			0			0
1503			Cu (%)			0.13			1.64			0.13			1.63			0			0			-0.01
1503			Fe (%)			32.7			0.96			32.7			0.96			0			0			0
1504			Ag (g/t)			252.69			1.33			252.69			1.33			0			0			0
1504			Au (g/t)			2.36			1.26			1.9			1.02			-0.46			-0.2			-0.25
1504			Pb (%)			2.32			1.52			2.32			1.52			0			0			0
1504			Zn (%)			3.79			1.46			3.79			1.46			0			0			0

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        432

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1504			Cu (%)			0.05			0.92			0.05			0.92			0			0			0
1504			Fe (%)			30.13			0.86			30.13			0.86			0			0			0
1505			Ag (g/t)			102.06			1.06			102.06			1.06			0			0			0
1505			Au (g/t)			0.75			1.26			0.75			1.26			0			0			0
1505			Pb (%)			0.95			1.44			0.95			1.44			0			0			0
1505			Zn (%)			1.73			1.26			1.73			1.26			0			0			0
1505			Cu (%)			0.04			1.28			0.04			1.24			0			0.02			-0.04
1505			Fe (%)			19.15			1.13			19.15			1.13			0			0			0
1601			Ag (g/t)			21.82			1.09			21.82			1.09			0			0			0
1601			Au (g/t)			0.16			0.71			0.16			0.71			0			0			0
1601			Pb (%)			1.28			1.46			1.28			1.46			0			0			0
1601			Zn (%)			1.81			0.76			1.81			0.76			0			0			0
1601			Cu (%)			0.09			0.98			0.09			0.98			0			0			0
1601			Fe (%)			13.68			0.55			13.68			0.55			0			0			0
1614			Ag (g/t)			14.43			0.88			14.43			0.88			0			0			0
1614			Au (g/t)			0.47			0.91			0.47			0.91			0			0			0
1614			Pb (%)			0.1			0.78			0.1			0.78			0			0			0
1614			Zn (%)			0.09			0.69			0.09			0.69			0			0			0
1614			Cu (%)			0.04			0.99			0.04			0.98			0			0			0
1614			Fe (%)			12.88			0.76			12.88			0.76			0			0			0
1631			Ag (g/t)			87.23			0.81			87.23			0.81			0			0			0
1631			Au (g/t)			1.74			0.98			1.74			0.98			0			0			0
1631			Pb (%)			1.67			0.8			1.67			0.8			0			0			0
1631			Zn (%)			2.16			1.17			2.16			1.17			0			0			0
1631			Cu (%)			0.06			0.77			0.06			0.75			0			0.02			-0.02
1631			Fe (%)			35.63			0.68			35.63			0.68			0			0			0
1632			Ag (g/t)			77.03			1.55			64.12			1.28			-12.91			-0.17			-0.27
1632			Au (g/t)			3.18			1.39			2.81			1.25			-0.38			-0.12			-0.14
1632			Pb (%)			1.44			1.94			0.82			0.77			-0.62			-0.43			-1.18
1632			Zn (%)			1.97			0.67			1.97			0.67			0			0			0
1632			Cu (%)			0.13			1.4			0.12			1.22			-0.02			-0.11			-0.18
1632			Fe (%)			33.16			0.99			33.16			0.99			0			0			0
2002			Ag (g/t)			72.51			1.87			72.51			1.87			0			0			0
2002			Au (g/t)			3.65			1.97			3.65			1.97			0			0			0
2002			Pb (%)			2.27			2.03			2.27			2.03			0			0			0
2002			Zn (%)			3.34			2.05			3.34			2.05			0			0			0
2002			Cu (%)			0.08			1.35			0.08			1.34			0			0.01			-0.02
2002			Fe (%)			31.48			1.01			31.48			1.01			0			0			0
2004			Ag (g/t)			186.38			1.23			186.38			1.23			0			0			0
2004			Au (g/t)			3.2			1.02			3.2			1.02			0			0			0
2004			Pb (%)			1.23			1.32			1.23			1.32			0			0			0
2004			Zn (%)			4.94			1.02			4.94			1.02			0			0			0
2004			Cu (%)			0.3			1.28			0.31			1.27			0			0			-0.01
2004			Fe (%)			43.24			0.72			43.24			0.72			0			0			0
2005			Ag (g/t)			19.57			0.85			19.57			0.85			0			0			0
2005			Au (g/t)			0.26			1.68			0.26			1.68			0			0			0
2005			Pb (%)			0.41			1.48			0.41			1.48			0			0			0
2005			Zn (%)			1.62			1.18			1.62			1.18			0			0			0
2005			Cu (%)			0.01			0.63			0.01			0.62			0			0.03			-0.01
2005			Fe (%)			13.43			0.58			13.43			0.58			0			0			0
2006			Ag (g/t)			32.99			0.89			32.99			0.89			0			0			0

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        433

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2006			Au (g/t)			0.62			2.32			0.62			2.32			0			0			0
2006			Pb (%)			2.81			1.9			1.86			1.52			-0.95			-0.34			-0.39
2006			Zn (%)			1.74			1.09			1.74			1.09			0			0			0
2006			Cu (%)			0.03			1.3			0.03			1.28			0			0.01			-0.01
2006			Fe (%)			12.1			1.05			12.1			1.05			0			0			0
2009			Ag (g/t)			69.51			0.79			69.51			0.79			0			0			0
2009			Au (g/t)			2.71			1.04			2.71			1.04			0			0			0
2009			Pb (%)			1.69			1.41			1.69			1.41			0			0			0
2009			Zn (%)			2.63			0.63			2.63			0.63			0			0			0
2009			Cu (%)			0.1			1.1			0.1			1.09			0			0.01			-0.01
2009			Fe (%)			42.65			0.8			42.65			0.8			0			0			0
2012			Ag (g/t)			74.49			1.04			71.77			1.04			-2.73			-0.04			0
2012			Au (g/t)			1.43			1.43			1.43			1.43			0			0			0
2012			Pb (%)			2.26			1.41			2.26			1.41			0			0			0
2012			Zn (%)			6.95			1.41			6.95			1.41			0			0			0
2012			Cu (%)			0.08			1.91			0.08			1.89			0			0.01			-0.01
2012			Fe (%)			25.01			0.84			25.01			0.84			0			0			0
2018			Ag (g/t)			176.5			1.01			176.5			1.01			0			0			0
2018			Au (g/t)			5.06			1.19			5.06			1.19			0			0			0
2018			Pb (%)			1.69			0.76			1.69			0.76			0			0			0
2018			Zn (%)			9.29			0.8			8.88			0.8			-0.41			-0.04			0
2018			Cu (%)			0.1			1.29			0.1			1.28			0			0.01			-0.01
2018			Fe (%)			53.38			1.03			53.38			1.03			0			0			0
2019			Ag (g/t)			159.32			0.73			159.32			0.73			0			0			0
2019			Au (g/t)			4.63			1.02			4.63			1.02			0			0			0
2019			Pb (%)			0.51			0.6			0.51			0.6			0			0			0
2019			Zn (%)			1.43			0.98			1.43			0.98			0			0			0
2019			Cu (%)			0.35			0.7			0.35			0.7			0			0			0
2019			Fe (%)			50.84			1.05			50.84			1.05			0			0			0
2020			Ag (g/t)			123.27			1.16			123.27			1.16			0			0			0
2020			Au (g/t)			13.42			1.38			10.7			1.26			-2.72			-0.2			-0.12
2020			Pb (%)			1.34			1.94			1.34			1.94			0			0			0
2020			Zn (%)			2.78			1.77			2.78			1.77			0			0			0
2020			Cu (%)			0.19			0.91			0.19			0.9			0			0			0
2020			Fe (%)			53.1			0.8			53.1			0.8			0			0			0
2026			Ag (g/t)			18.6			0.86			18.6			0.86			0			0			0
2026			Au (g/t)			0.77			0.99			0.77			0.99			0			0			0
2026			Pb (%)			0.89			1.5			0.89			1.5			0			0			0
2026			Zn (%)			2.17			1.24			2.17			1.24			0			0			0
2026			Cu (%)			0.02			1.17			0.02			0.97			0			0.2			-0.2
2026			Fe (%)			16.5			0.69			16.5			0.69			0			0			0
2091			Ag (g/t)			246.09			1.45			221.43			1.22			-24.66			-0.1			-0.23
2091			Au (g/t)			11.33			1.93			10.15			1.5			-1.18			-0.1			-0.42
2091			Pb (%)			2.03			1.21			2			1.17			-0.03			-0.01			-0.05
2091			Zn (%)			7.39			1.34			7.11			1.21			-0.28			-0.04			-0.14
2091			Cu (%)			0.39			1.71			0.37			1.51			-0.02			-0.04			-0.2
2091			Fe (%)			74.85			0.89			73.71			0.86			-1.13			-0.02			-0.03
2101			Ag (g/t)			40.64			1.25			40.64			1.25			0			0			0
2101			Au (g/t)			1.49			1.77			1.49			1.77			0			0			0
2101			Pb (%)			0.41			0.79			0.41			0.79			0			0			0
2101			Zn (%)			2.3			1.18			2.13			1.22			-0.16			-0.07			0.05

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        434

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2101			Cu (%)			0.05			1.76			0.05			1.75			0			0.01			-0.02
2101			Fe (%)			12.82			1.12			12.82			1.12			0			0			0
2103			Ag (g/t)			69.05			1.46			69.05			1.46			0			0			0
2103			Au (g/t)			2.47			2.25			2.47			2.25			0			0			0
2103			Pb (%)			1.48			1.72			1.48			1.72			0			0			0
2103			Zn (%)			5.09			1.7			5.09			1.7			0			0			0
2103			Cu (%)			0.1			1.75			0.11			1.74			0			0			-0.01
2103			Fe (%)			30.74			1.48			30.74			1.48			0			0			0
2104			Ag (g/t)			99.57			1.71			98.01			1.68			-1.56			-0.02			-0.03
2104			Au (g/t)			3.76			1.64			3.52			1.51			-0.24			-0.06			-0.13
2104			Pb (%)			1.36			1.32			1.32			1.24			-0.05			-0.03			-0.08
2104			Zn (%)			4.24			1.28			4.08			1.21			-0.16			-0.04			-0.08
2104			Cu (%)			0.18			1.63			0.18			1.54			-0.01			-0.03			-0.08
2104			Fe (%)			37.59			1.21			37.59			1.21			0			0			0
2106			Ag (g/t)			47.41			0.87			47.41			0.87			0			0			0
2106			Au (g/t)			2.82			1.35			2.82			1.35			0			0			0
2106			Pb (%)			0.94			1			0.94			1			0			0			0
2106			Zn (%)			2.91			1.18			2.91			1.18			0			0			0
2106			Cu (%)			0.2			2.4			0.2			2.39			0			0			0
2106			Fe (%)			25.86			0.68			25.86			0.68			0			0			0
2113			Ag (g/t)			37.39			0.94			37.39			0.94			0			0			0
2113			Au (g/t)			0.94			1.03			0.94			1.03			0			0			0
2113			Pb (%)			0.92			0.7			0.92			0.7			0			0			0
2113			Zn (%)			1.76			0.92			1.76			0.92			0			0			0
2113			Cu (%)			0.08			0.68			0.08			0.68			0			0			0
2113			Fe (%)			17.82			0.87			17.82			0.87			0			0			0
2115			Ag (g/t)			25.11			0.62			25.11			0.62			0			0			0
2115			Au (g/t)			0.37			0.53			0.37			0.53			0			0			0
2115			Pb (%)			0.85			1.14			0.85			1.14			0			0			0
2115			Zn (%)			1.92			1.08			1.92			1.08			0			0			0
2115			Cu (%)			0.05			0.67			0.05			0.66			0			0.01			-0.01
2115			Fe (%)			16.47			0.86			16.47			0.86			0			0			0
2117			Ag (g/t)			140.7			1.26			140.7			1.26			0			0			0
2117			Au (g/t)			4.72			1.66			4.72			1.66			0			0			0
2117			Pb (%)			1.21			1.26			1.21			1.26			0			0			0
2117			Zn (%)			1.77			1.07			1.77			1.07			0			0			0
2117			Cu (%)			0.26			1.69			0.26			1.69			0			0			-0.01
2117			Fe (%)			34.61			1.01			34.61			1.01			0			0			0
2118			Ag (g/t)			58.32			0.72			58.32			0.72			0			0			0
2118			Au (g/t)			2.36			1.02			2.36			1.02			0			0			0
2118			Pb (%)			1.22			0.69			1.22			0.69			0			0			0
2118			Zn (%)			3.19			0.62			3.19			0.62			0			0			0
2118			Cu (%)			0.15			1.43			0.15			1.42			0			0			-0.01
2118			Fe (%)			38.73			1.04			38.73			1.04			0			0			0
2120			Ag (g/t)			169.33			0.54			169.33			0.54			0			0			0
2120			Au (g/t)			5.39			0.7			5.39			0.7			0			0			0
2120			Pb (%)			1.01			0.83			1.01			0.83			0			0			0
2120			Zn (%)			3.01			1.08			3.01			1.08			0			0			0
2120			Cu (%)			0.37			0.68			0.37			0.68			0			0			0
2120			Fe (%)			41.91			0.61			41.91			0.61			0			0			0
2122			Ag (g/t)			236.89			1.81			236.89			1.81			0			0			0

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        435

---

2122			Au (g/t)			1.65			1.28			1.65			1.28			0			0			0
2122			Pb (%)			1.31			1.48			1.31			1.48			0			0			0
2122			Zn (%)			0.68			1.2			0.68			1.2			0			0			0
2122			Cu (%)			0.08			1			0.08			0.98			0			0.01			-0.02
2122			Fe (%)			24.71			0.93			24.71			0.93			0			0			0
2127			Ag (g/t)			19.8			0.77			19.8			0.77			0			0			0
2127			Au (g/t)			0.07			0.83			0.07			0.83			0			0			0
2127			Pb (%)			0.12			0.69			0.12			0.69			0			0			0
2127			Zn (%)			1.27			0.95			1.27			0.95			0			0			0
2127			Cu (%)			0.1			0.89			0.1			0.89			0			0			0
2127			Fe (%)			9.84			0.3			9.84			0.3			0			0			0
2130			Ag (g/t)			19.55			0.64			19.55			0.64			0			0			0
2130			Au (g/t)			0.9			1.49			0.9			1.49			0			0			0
2130			Pb (%)			1.43			0.97			1.43			0.97			0			0			0
2130			Zn (%)			2.07			1.45			2.07			1.45			0			0			0
2130			Cu (%)			0.15			1.54			0.15			1.54			0			0			0
2130			Fe (%)			11.62			0.55			11.62			0.55			0			0			0
2131			Ag (g/t)			21.9			0.74			21.9			0.74			0			0			0
2131			Au (g/t)			1.28			0.9			1.28			0.9			0			0			0
2131			Pb (%)			0.69			1.14			0.69			1.14			0			0			0
2131			Zn (%)			1.53			1.34			1.53			1.34			0			0			0
2131			Cu (%)			0.03			0.74			0.03			0.74			0			0.01			0
2131			Fe (%)			23.49			0.61			23.49			0.61			0			0			0
2132			Ag (g/t)			22.7			1.36			22.7			1.36			0			0			0
2132			Au (g/t)			0.47			1.82			0.47			1.82			0			0			0
2132			Pb (%)			0.64			2.15			0.64			2.15			0			0			0
2132			Zn (%)			1.14			1.09			1.14			1.09			0			0			0
2132			Cu (%)			0.05			1.98			0.05			1.97			0			0			-0.01
2132			Fe (%)			11.08			1.22			11.08			1.22			0			0			0
2203			Ag (g/t)			126.03			1.48			105.84			1.21			-20.18			-0.16			-0.28
2203			Au (g/t)			0.65			1.63			0.65			1.63			0			0			0
2203			Pb (%)			1.46			0.76			1.46			0.76			0			0			0
2203			Zn (%)			5.1			1.11			5.1			1.11			0			0			0
2203			Cu (%)			0.02			0.62			0.02			0.7			0			0.13			0.07
2203			Fe (%)			20.8			0.82			20.8			0.82			0			0			0
2207			Ag (g/t)			131.52			1.62			131.52			1.62			0			0			0
2207			Au (g/t)			3.82			1.68			3.82			1.68			0			0			0
2207			Pb (%)			1.12			1.36			1.12			1.36			0			0			0
2207			Zn (%)			1.13			0.82			1.13			0.82			0			0			0
2207			Cu (%)			0.03			0.98			0.03			0.91			0			0.04			-0.07
2207			Fe (%)			19.12			0.67			19.12			0.67			0			0			0
2213			Ag (g/t)			83.75			0.96			83.75			0.96			0			0			0
2213			Au (g/t)			4.74			0.75			4.74			0.75			0			0			0
2213			Pb (%)			0.75			0.82			0.75			0.82			0			0			0
2213			Zn (%)			1.57			1.26			1.57			1.26			0			0			0
2213			Cu (%)			0.14			0.59			0.14			0.6			0			0.01			0.01
2213			Fe (%)			26.85			0.43			26.85			0.43			0			0			0
2215			Ag (g/t)			22.12			1.28			22.12			1.28			0			0			0
2215			Au (g/t)			0.43			1.07			0.43			1.07			0			0			0
2215			Pb (%)			1.22			1.2			1.22			1.2			0			0			0
2215			Zn (%)			16.14			1.49			16.14			1.49			0			0			0

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        436

---

2215			Cu (%)			0.05			1.18			0.05			1.18			0			0			0
2215			Fe (%)			39.98			0.88			39.98			0.88			0			0			0
2216			Ag (g/t)			28.53			0.82			28.53			0.82			0			0			0
2216			Au (g/t)			0.85			1.37			0.85			1.37			0			0			0
2216			Pb (%)			0.51			1.17			0.51			1.17			0			0			0
2216			Zn (%)			1.96			1.62			1.96			1.62			0			0			0
2216			Cu (%)			0.02			0.76			0.02			0.71			0			0.03			-0.04
2216			Fe (%)			13.51			0.47			13.51			0.47			0			0			0
2219			Ag (g/t)			159.53			1.36			159.53			1.36			0			0			0
2219			Au (g/t)			5.8			1.34			5.8			1.34			0			0			0
2219			Pb (%)			1.72			0.71			1.72			0.71			0			0			0
2219			Zn (%)			1.76			1.63			1.76			1.63			0			0			0
2219			Cu (%)			0.25			1.3			0.25			1.3			0			0			0
2219			Fe (%)			46.91			1.02			46.91			1.02			0			0			0
2221			Ag (g/t)			34.16			0.75			34.16			0.75			0			0			0
2221			Au (g/t)			0.69			1.37			0.69			1.37			0			0			0
2221			Pb (%)			1.05			0.88			1.05			0.88			0			0			0
2221			Zn (%)			1.07			1.01			1.07			1.01			0			0			0
2221			Cu (%)			0.03			1.17			0.03			1.15			0			0.01			-0.02
2221			Fe (%)			14.05			0.89			14.05			0.89			0			0			0
2222			Ag (g/t)			35.97			0.84			35.97			0.84			0			0			0
2222			Au (g/t)			0.31			0.48			0.31			0.48			0			0			0
2222			Pb (%)			1.84			1.18			1.84			1.18			0			0			0
2222			Zn (%)			20.38			1.39			16.46			1.28			-3.91			-0.19			-0.11
2222			Cu (%)			0.15			1.45			0.15			1.43			0			0.01			-0.02
2222			Fe (%)			12.75			0.45			12.75			0.45			0			0			0
2223			Ag (g/t)			105.52			1.35			95.32			1.25			-10.2			-0.1			-0.1
2223			Au (g/t)			2.77			1.56			2.77			1.56			0			0			0
2223			Pb (%)			1.15			0.92			1.15			0.92			0			0			0
2223			Zn (%)			3.84			1.17			3.84			1.17			0			0			0
2223			Cu (%)			0.04			0.87			0.04			0.85			0			0.05			-0.02
2223			Fe (%)			27.96			1.22			27.96			1.22			0			0			0
2224			Ag (g/t)			46.72			0.32			46.72			0.32			0			0			0
2224			Au (g/t)			1.41			1			1.41			1			0			0			0
2224			Pb (%)			5.15			0.91			2.6			0.81			-2.55			-0.5			-0.09
2224			Zn (%)			1.23			0.84			1.23			0.84			0			0			0
2224			Cu (%)			0.03			0.24			0.04			0.17			0			0.06			-0.07
2224			Fe (%)			12.9			0.78			12.9			0.78			0			0			0
2225			Ag (g/t)			44.93			1.47			44.93			1.47			0			0			0
2225			Au (g/t)			3.86			1.31			3.86			1.31			0			0			0
2225			Pb (%)			0.39			1.38			0.39			1.38			0			0			0
2225			Zn (%)			0.49			1.11			0.49			1.11			0			0			0
2225			Cu (%)			0.1			1.47			0.1			1.47			0			0			0
2225			Fe (%)			26.53			1.22			26.53			1.22			0			0			0
2235			Ag (g/t)			19.09			1.11			19.09			1.11			0			0			0
2235			Au (g/t)			4.07			1.52			4.07			1.52			0			0			0
2235			Pb (%)			0.56			1.22			0.56			1.22			0			0			0
2235			Zn (%)			3.28			0.78			3.28			0.78			0			0			0
2235			Cu (%)			0.03			0.58			0.03			0.58			0			0			0
2235			Fe (%)			14.27			1.03			14.27			1.03			0			0			0
2236			Ag (g/t)			106.32			0.68			106.32			0.68			0			0			0

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        437

---

2236			Au (g/t)			1.03			0.79			1.03			0.79			0			0			0
2236			Pb (%)			0.85			0.59			0.85			0.59			0			0			0
2236			Zn (%)			2.29			0.69			2.29			0.69			0			0			0
2236			Cu (%)			0.04			0.8			0.04			0.79			0			0.01			-0.02
2236			Fe (%)			27.33			0.37			27.33			0.37			0			0			0
2309			Ag (g/t)			117.84			1.65			117.84			1.65			0			0			0
2309			Au (g/t)			1.65			1.51			1.65			1.51			0			0			0
2309			Pb (%)			1.47			1.43			1.47			1.43			0			0			0
2309			Zn (%)			0.62			0.98			0.62			0.98			0			0			0
2309			Cu (%)			0.11			1.15			0.11			1.15			0			0			0
2309			Fe (%)			23.55			0.61			23.55			0.61			0			0			0
2311			Ag (g/t)			46.16			0.78			46.16			0.78			0			0			0
2311			Au (g/t)			1.72			0.85			1.72			0.85			0			0			0
2311			Pb (%)			1.07			1.72			1.07			1.72			0			0			0
2311			Zn (%)			3.22			1.39			3.22			1.39			0			0			0
2311			Cu (%)			0.07			1.22			0.07			1.21			0			0.01			-0.01
2311			Fe (%)			24.58			0.61			24.58			0.61			0			0			0
2314			Ag (g/t)			74.95			1.27			74.95			1.27			0			0			0
2314			Au (g/t)			0.68			0.57			0.68			0.57			0			0			0
2314			Pb (%)			1.06			0.78			1.06			0.78			0			0			0
2314			Zn (%)			2.82			1.34			2.82			1.34			0			0			0
2314			Cu (%)			0.16			0.74			0.16			0.74			0			0			0
2314			Fe (%)			28			0.65			28			0.65			0			0			0
2327			Ag (g/t)			33.13			0.55			33.13			0.55			0			0			0
2327			Au (g/t)			0.88			0.99			0.88			0.99			0			0			0
2327			Pb (%)			0.24			0.29			0.24			0.29			0			0			0
2327			Zn (%)			1.03			0.67			1.03			0.67			0			0			0
2327			Cu (%)			0.02			0.77			0.02			0.77			0			0.01			0
2327			Fe (%)			16.08			0.75			16.08			0.75			0			0			0
2400			Ag (g/t)			144.66			0.89			142.39			0.89			-2.27			-0.02			0
2400			Au (g/t)			2.76			1			2.76			1			0			0			0
2400			Pb (%)			1.34			2.11			1.11			1.68			-0.24			-0.17			-0.43
2400			Zn (%)			1.85			1.41			1.85			1.41			0			0			0
2400			Cu (%)			0.12			1.47			0.12			1.45			0			0.01			-0.02
2400			Fe (%)			41.14			0.7			41.14			0.7			0			0			0
2401			Ag (g/t)			43.93			1.02			43.93			1.02			0			0			0
2401			Au (g/t)			0.44			2.36			0.44			2.36			0			0			0
2401			Pb (%)			2.58			1.68			2.55			1.68			-0.03			-0.01			0
2401			Zn (%)			1.87			1.48			1.87			1.48			0			0			0
2401			Cu (%)			0.05			1.64			0.05			1.63			0			0			-0.01
2401			Fe (%)			15.4			1.68			15.4			1.68			0			0			0
2405			Ag (g/t)			134.01			1.43			51.73			0.64			-82.28			-0.61			-0.79
2405			Au (g/t)			1.81			1.58			0.56			0.79			-1.25			-0.69			-0.79
2405			Pb (%)			1.6			0.86			1.6			0.86			0			0			0
2405			Zn (%)			3.33			1.16			2.27			0.94			-1.06			-0.32			-0.22
2405			Cu (%)			0.05			0.58			0.06			0.6			0			0.04			0.01
2405			Fe (%)			55.31			1.12			48.07			0.99			-7.24			-0.13			-0.12
2408			Ag (g/t)			69.81			1.17			69.81			1.17			0			0			0
2408			Au (g/t)			1.03			0.66			1.03			0.66			0			0			0
2408			Pb (%)			0.47			1.41			0.47			1.41			0			0			0
2408			Zn (%)			0.69			1.38			0.69			1.38			0			0			0

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        438

---

2408			Cu (%)			0.09			1.75			0.09			1.75			0			0			0
2408			Fe (%)			13.43			0.54			13.43			0.54			0			0			0
2410			Ag (g/t)			117.02			1.79			119.56			1.8			2.54			0.02			0.01
2410			Au (g/t)			2.39			2.03			2.43			2.03			0.04			0.02			0
2410			Pb (%)			1.4			1.23			1.4			1.23			0			0			0
2410			Zn (%)			3.33			1.07			3.36			1.07			0.03			0.01			0
2410			Cu (%)			0.08			1.98			0.08			1.96			0			0.01			-0.02
2410			Fe (%)			30.62			1.23			30.62			1.23			0			0			0
2411			Ag (g/t)			29.57			0.81			29.57			0.81			0			0			0
2411			Au (g/t)			2.3			1.34			2.3			1.34			0			0			0
2411			Pb (%)			0.34			0.84			0.34			0.84			0			0			0
2411			Zn (%)			0.75			0.81			0.75			0.81			0			0			0
2411			Cu (%)			0.05			0.86			0.05			0.86			0			0			0
2411			Fe (%)			34.54			1.13			34.54			1.13			0			0			0
2417			Ag (g/t)			33.86			0.84			33.86			0.84			0			0			0
2417			Au (g/t)			1.34			1.01			1.34			1.01			0			0			0
2417			Pb (%)			0.33			0.99			0.33			0.99			0			0			0
2417			Zn (%)			0.59			1.07			0.59			1.07			0			0			0
2417			Cu (%)			0.01			0.35			0.01			0.35			0			0			0
2417			Fe (%)			20.67			0.61			20.67			0.61			0			0			0
2424			Ag (g/t)			21.66			0.42			21.66			0.42			0			0			0
2424			Au (g/t)			0.59			0.84			0.59			0.84			0			0			0
2424			Pb (%)			0.59			0.8			0.59			0.8			0			0			0
2424			Zn (%)			1.44			0.78			1.44			0.78			0			0			0
2424			Cu (%)			0.02			0.89			0.02			0.89			0			0			0
2424			Fe (%)			15.1			0.51			15.1			0.51			0			0			0
2425			Ag (g/t)			26.48			0.77			26.48			0.77			0			0			0
2425			Au (g/t)			0.2			1.66			0.2			1.66			0			0			0
2425			Pb (%)			1.73			1.26			1.65			1.24			-0.08			-0.05			-0.02
2425			Zn (%)			0.92			1.11			0.92			1.11			0			0			0
2425			Cu (%)			0.21			1.08			0.21			1.08			0			0			0
2425			Fe (%)			2.2			0.74			2.2			0.74			0			0			0
2502			Ag (g/t)			42.23			1.12			42.23			1.12			0			0			0
2502			Au (g/t)			2.15			1.39			2.15			1.39			0			0			0
2502			Pb (%)			1.61			1.64			1.47			1.6			-0.15			-0.09			-0.04
2502			Zn (%)			1.45			1.36			1.45			1.36			0			0			0
2502			Cu (%)			0.2			1.44			0.2			1.44			0			0			0
2502			Fe (%)			34.42			1.01			34.42			1.01			0			0			0
2507			Ag (g/t)			79.07			1.57			79.07			1.57			0			0			0
2507			Au (g/t)			2.82			2.87			2.82			2.87			0			0			0
2507			Pb (%)			0.88			1.5			0.88			1.5			0			0			0
2507			Zn (%)			2.19			1.56			2.19			1.56			0			0			0
2507			Cu (%)			0.12			1.67			0.12			1.65			0			0.01			-0.02
2507			Fe (%)			21.73			1.39			21.73			1.39			0			0			0
2510			Ag (g/t)			48.26			0.83			48.26			0.83			0			0			0
2510			Au (g/t)			1.28			1.24			1.28			1.24			0			0			0
2510			Pb (%)			0.71			0.83			0.71			0.83			0			0			0
2510			Zn (%)			0.89			0.84			0.89			0.84			0			0			0
2510			Cu (%)			0.24			1.26			0.24			1.26			0			0			0
2510			Fe (%)			53.06			1.33			46.81			1.2			-6.25			-0.12			-0.13
2538			Ag (g/t)			12.49			0.73			12.49			0.73			0			0			0

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        439

---

2538			Au (g/t)			0.1			0.97			0.1			0.97			0			0			0
2538			Pb (%)			0.8			1.47			0.8			1.47			0			0			0
2538			Zn (%)			3.1			1.97			2.85			1.9			-0.25			-0.08			-0.07
2538			Cu (%)			0.02			1.5			0.02			1.44			0			0.03			-0.06
2538			Fe (%)			6.19			0.55			6.19			0.55			0			0			0
2614			Ag (g/t)			51.77			0.59			51.77			0.59			0			0			0
2614			Au (g/t)			0.4			0.83			0.4			0.83			0			0			0
2614			Pb (%)			0.96			0.72			0.96			0.72			0			0			0
2614			Zn (%)			3.33			1.04			3.33			1.04			0			0			0
2614			Cu (%)			0.27			1.67			0.27			1.66			0			0.01			-0.01
2614			Fe (%)			35.15			0.53			35.15			0.53			0			0			0
2729			Ag (g/t)			36.32			0.72			36.32			0.72			0			0			0
2729			Au (g/t)			1.87			0.82			1.87			0.82			0			0			0
2729			Pb (%)			0.51			0.28			0.51			0.28			0			0			0
2729			Zn (%)			0.61			0.51			0.61			0.51			0			0			0
2729			Cu (%)			0.09			1.06			0.09			1.06			0			0			0
2729			Fe (%)			32.21			0.72			32.21			0.72			0			0			0
2802			Ag (g/t)			63.43			1.1			62.05			1.08			-1.38			-0.02			-0.02
2802			Au (g/t)			1.36			1.32			1.36			1.32			0			0			0
2802			Pb (%)			2.56			1.12			2.56			1.12			0			0			0
2802			Zn (%)			4.83			1.39			4.83			1.39			0			0			0
2802			Cu (%)			0.05			1.08			0.05			1.06			0			0.01			-0.02
2802			Fe (%)			16.78			1.07			16.78			1.07			0			0			0
3001			Ag (g/t)			68.96			2.07			49.24			1.29			-19.72			-0.29			-0.78
3001			Au (g/t)			1.34			1.49			1.34			1.49			0			0			0
3001			Pb (%)			0.64			0.97			0.64			0.97			0			0			0
3001			Zn (%)			1.62			1.21			1.62			1.21			0			0			0
3001			Cu (%)			0.1			1.67			0.1			1.67			0			0			0
3001			Fe (%)			28.67			1.16			28.67			1.16			0			0			0
3002			Ag (g/t)			28.9			0.54			28.9			0.54			0			0			0
3002			Au (g/t)			0.72			0.52			0.72			0.52			0			0			0
3002			Pb (%)			0.76			1.33			0.76			1.33			0			0			0
3002			Zn (%)			1.5			0.82			1.5			0.82			0			0			0
3002			Cu (%)			0.02			0.5			0.03			0.5			0			0.01			0
3002			Fe (%)			14.06			0.45			14.06			0.45			0			0			0
3003			Ag (g/t)			17.77			0.93			17.77			0.93			0			0			0
3003			Au (g/t)			0.19			0.85			0.19			0.85			0			0			0
3003			Pb (%)			0.39			0.84			0.39			0.84			0			0			0
3003			Zn (%)			0.78			1.05			0.78			1.05			0			0			0
3003			Cu (%)			0.04			0.56			0.04			0.56			0			0			0
3003			Fe (%)			16.71			0.67			16.71			0.67			0			0			0
3004			Ag (g/t)			31.97			0.93			31.97			0.93			0			0			0
3004			Au (g/t)			1.19			1.29			1.19			1.29			0			0			0
3004			Pb (%)			0.85			1.79			0.85			1.79			0			0			0
3004			Zn (%)			1.34			1.09			1.34			1.09			0			0			0
3004			Cu (%)			0.05			1.37			0.05			1.38			0			0.02			0.02
3004			Fe (%)			16.74			0.77			16.74			0.77			0			0			0
3005			Ag (g/t)			61.28			0.68			61.28			0.68			0			0			0
3005			Au (g/t)			2.51			1.2			2.51			1.2			0			0			0
3005			Pb (%)			0.83			0.7			0.83			0.7			0			0			0
3005			Zn (%)			1.91			1.29			1.91			1.29			0			0			0

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        440

---

3005			Cu (%)			0.05			0.75			0.05			0.74			0			0.02			-0.01
3005			Fe (%)			35.41			0.76			35.41			0.76			0			0			0
3006			Ag (g/t)			55.43			1.18			55.43			1.18			0			0			0
3006			Au (g/t)			1.18			0.98			1.18			0.98			0			0			0
3006			Pb (%)			0.85			0.92			0.85			0.92			0			0			0
3006			Zn (%)			5.2			0.97			4.66			0.9			-0.53			-0.1			-0.07
3006			Cu (%)			0.15			1.37			0.15			1.37			0			0			0
3006			Fe (%)			35.87			1.03			35.87			1.03			0			0			0
3007			Ag (g/t)			56.4			1.07			56.4			1.07			0			0			0
3007			Au (g/t)			0.49			1.48			0.49			1.48			0			0			0
3007			Pb (%)			0.9			1.21			0.9			1.21			0			0			0
3007			Zn (%)			2			0.44			2			0.44			0			0			0
3007			Cu (%)			0.12			1.06			0.12			1.06			0			0			0
3007			Fe (%)			27.03			0.55			27.03			0.55			0			0			0
3008			Ag (g/t)			83.32			0.85			83.32			0.85			0			0			0
3008			Au (g/t)			0.32			0.53			0.32			0.53			0			0			0
3008			Pb (%)			0.56			0.61			0.56			0.61			0			0			0
3008			Zn (%)			1.94			0.45			1.94			0.45			0			0			0
3008			Cu (%)			0.04			0.55			0.04			0.55			0			0			0
3008			Fe (%)			40.74			0.97			40.74			0.97			0			0			0
3009			Ag (g/t)			90.82			1.49			90.82			1.49			0			0			0
3009			Au (g/t)			6.23			2.56			6.23			2.56			0			0			0
3009			Pb (%)			1.2			1.17			1.2			1.17			0			0			0
3009			Zn (%)			1.71			0.9			1.71			0.9			0			0			0
3009			Cu (%)			0.12			1.66			0.12			1.65			0			0.01			0
3009			Fe (%)			57.44			0.99			57.44			0.99			0			0			0
3010			Ag (g/t)			29.98			1.27			29.98			1.27			0			0			0
3010			Au (g/t)			2.12			1.01			2.12			1.01			0			0			0
3010			Pb (%)			0.5			1.09			0.5			1.09			0			0			0
3010			Zn (%)			0.61			0.94			0.61			0.94			0			0			0
3010			Cu (%)			0.1			1.63			0.1			1.63			0			0			0
3010			Fe (%)			18.53			0.77			18.53			0.77			0			0			0
3011			Ag (g/t)			20.77			0.66			20.77			0.66			0			0			0
3011			Au (g/t)			0.64			0.59			0.64			0.59			0			0			0
3011			Pb (%)			0.66			0.95			0.66			0.95			0			0			0
3011			Zn (%)			2.56			0.74			2.55			0.74			-0.01			-0.01			0
3011			Cu (%)			0.03			0.77			0.03			0.76			0			0.01			-0.01
3011			Fe (%)			17.48			0.77			16.37			0.67			-1.1			-0.06			-0.1
3100			Ag (g/t)			13.19			0.73			13.19			0.73			0			0			0
3100			Au (g/t)			0.11			0.34			0.11			0.34			0			0			0
3100			Pb (%)			0.48			0.75			0.48			0.75			0			0			0
3100			Zn (%)			3.46			0.89			3.34			0.89			-0.13			-0.04			0
3100			Cu (%)			0.01			0.52			0.01			0.59			0			0.17			0.07
3100			Fe (%)			4.96			0.67			4.96			0.67			0			0			0
3101			Ag (g/t)			58.14			1.34			54.41			1.14			-3.73			-0.06			-0.2
3101			Au (g/t)			2.26			1.31			2.18			1.2			-0.08			-0.04			-0.1
3101			Pb (%)			0.64			1.66			0.61			1.57			-0.02			-0.03			-0.08
3101			Zn (%)			1.58			1.32			1.49			1.19			-0.09			-0.06			-0.13
3101			Cu (%)			0.14			1.99			0.12			1.63			-0.01			-0.11			-0.36
3101			Fe (%)			30.86			1.08			30.86			1.08			0			0			0
3102			Ag (g/t)			107.27			1.43			99.83			1.32			-7.43			-0.07			-0.11

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        441

---

3102			Au (g/t)			1.94			1.07			1.94			1.07			0			0			0
3102			Pb (%)			1.06			1.57			1.06			1.57			0			0			0
3102			Zn (%)			1.91			1.24			1.9			1.23			-0.01			0			-0.01
3102			Cu (%)			0.11			1.24			0.11			1.24			0			0			0
3102			Fe (%)			31.26			0.75			31.26			0.75			0			0			0
3103			Ag (g/t)			59.16			1.56			52.54			1.34			-6.62			-0.11			-0.21
3103			Au (g/t)			3.32			1.53			2.56			1.33			-0.77			-0.23			-0.2
3103			Pb (%)			0.35			0.86			0.35			0.86			0			0			0
3103			Zn (%)			1.28			1.89			0.84			0.88			-0.44			-0.34			-1.02
3103			Cu (%)			0.18			2.35			0.16			2.2			-0.02			-0.14			-0.16
3103			Fe (%)			34.74			1.2			34.74			1.2			0			0			0
3121			Ag (g/t)			32.97			0.8			32.97			0.8			0			0			0
3121			Au (g/t)			1.64			1.55			1.64			1.55			0			0			0
3121			Pb (%)			0.36			0.96			0.36			0.96			0			0			0
3121			Zn (%)			0.87			0.95			0.87			0.95			0			0			0
3121			Cu (%)			0.11			1.33			0.11			1.33			0			0			-0.01
3121			Fe (%)			25.84			1.17			25.05			1.11			-0.79			-0.03			-0.05
3191			Ag (g/t)			74.14			1.43			74.14			1.43			0			0			0
3191			Au (g/t)			3.79			1.49			3.79			1.49			0			0			0
3191			Pb (%)			0.54			1.38			0.54			1.38			0			0			0
3191			Zn (%)			1.97			1.65			1.97			1.65			0			0			0
3191			Cu (%)			0.15			1.93			0.15			1.93			0			0			-0.01
3191			Fe (%)			35.56			1.1			35.56			1.1			0			0			0
3192			Ag (g/t)			65.74			1.15			65.74			1.15			0			0			0
3192			Au (g/t)			2.24			0.95			2.24			0.95			0			0			0
3192			Pb (%)			0.59			1.55			0.59			1.55			0			0			0
3192			Zn (%)			1.57			0.96			1.57			0.96			0			0			0
3192			Cu (%)			0.11			1.34			0.11			1.34			0			0			0
3192			Fe (%)			31.29			0.63			31.29			0.63			0			0			0
3193			Ag (g/t)			35.42			1.39			35.42			1.39			0			0			0
3193			Au (g/t)			2.38			2.27			2.38			2.27			0			0			0
3193			Pb (%)			0.34			0.81			0.34			0.81			0			0			0
3193			Zn (%)			1			1.02			1			1.02			0			0			0
3193			Cu (%)			0.05			1.19			0.05			1.18			0			0.01			-0.01
3193			Fe (%)			24.02			1.08			24.02			1.08			0			0			0
3204			Ag (g/t)			41.23			0.82			41.23			0.82			0			0			0
3204			Au (g/t)			1.18			1.06			1.18			1.06			0			0			0
3204			Pb (%)			0.36			1.36			0.36			1.36			0			0			0
3204			Zn (%)			1.01			1.56			1.01			1.56			0			0			0
3204			Cu (%)			0.07			1.79			0.07			1.78			0			0.01			-0.01
3204			Fe (%)			18.04			0.82			18.04			0.82			0			0			0
3205			Ag (g/t)			34.21			2.66			34.21			2.66			0			0			0
3205			Au (g/t)			0.23			1.2			0.23			1.2			0			0			0
3205			Pb (%)			0.31			1.04			0.31			1.04			0			0			0
3205			Zn (%)			0.98			0.83			0.98			0.83			0			0			0
3205			Cu (%)			0.03			0.83			0.03			0.83			0			0.01			0
3205			Fe (%)			13.45			0.82			13.45			0.82			0			0			0
3206			Ag (g/t)			32.28			1.01			32.28			1.01			0			0			0
3206			Au (g/t)			1.03			1.12			1.03			1.12			0			0			0
3206			Pb (%)			0.49			1.34			0.49			1.34			0			0			0
3206			Zn (%)			1.73			1.5			1.73			1.5			0			0			0

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        442

---

3206			Cu (%)			0.07			0.99			0.07			1			0			0.01			0.01
3206			Fe (%)			23.73			0.74			23.73			0.74			0			0			0
3207			Ag (g/t)			10.94			0.44			10.94			0.44			0			0			0
3207			Au (g/t)			0.59			1			0.59			1			0			0			0
3207			Pb (%)			0.22			1.31			0.22			1.31			0			0			0
3207			Zn (%)			0.45			1.05			0.45			1.05			0			0			0
3207			Cu (%)			0.05			0.88			0.05			0.88			0			0			0
3207			Fe (%)			15.13			0.53			15.13			0.53			0			0			0
3222			Ag (g/t)			20.1			1.2			20.1			1.2			0			0			0
3222			Au (g/t)			0.62			2.99			0.54			2.8			-0.08			-0.13			-0.19
3222			Pb (%)			0.57			1.53			0.57			1.53			0			0			0
3222			Zn (%)			1.16			1.12			1.16			1.12			0			0			0
3222			Cu (%)			0.07			1.8			0.07			1.8			0			0			0
3222			Fe (%)			11.72			1.13			11.72			1.13			0			0			0
3223			Ag (g/t)			31.2			1.36			31.2			1.36			0			0			0
3223			Au (g/t)			1.03			1.74			1.03			1.74			0			0			0
3223			Pb (%)			1.33			1.56			1.33			1.56			0			0			0
3223			Zn (%)			2.4			1.57			2.38			1.56			-0.01			-0.01			0
3223			Cu (%)			0.07			2.04			0.07			2.02			0			0.01			-0.02
3223			Fe (%)			18			1.1			18			1.1			0			0			0
3224			Ag (g/t)			20.79			1.25			17.74			0.92			-3.05			-0.15			-0.33
3224			Au (g/t)			0.37			1.68			0.37			1.68			0			0			0
3224			Pb (%)			0.43			1.02			0.43			1.02			0			0			0
3224			Zn (%)			1.08			0.77			1.08			0.77			0			0			0
3224			Cu (%)			0.1			2.39			0.1			2.39			0			0			0
3224			Fe (%)			11.04			0.38			11.04			0.38			0			0			0
3225			Ag (g/t)			51.64			0.61			51.64			0.61			0			0			0
3225			Au (g/t)			2.2			0.84			2.2			0.84			0			0			0
3225			Pb (%)			3.52			0.85			3.52			0.85			0			0			0
3225			Zn (%)			2.44			0.54			2.44			0.54			0			0			0
3225			Cu (%)			0.29			0.6			0.29			0.6			0			0			0
3225			Fe (%)			30.78			0.56			30.78			0.56			0			0			0
3226			Ag (g/t)			29.25			1.6			29.25			1.6			0			0			0
3226			Au (g/t)			0.69			1.9			0.62			1.75			-0.07			-0.1			-0.15
3226			Pb (%)			0.96			1.09			0.96			1.09			0			0			0
3226			Zn (%)			1.87			1.56			1.82			1.58			-0.05			-0.03			0.01
3226			Cu (%)			0.11			1.5			0.11			1.5			0			0			0
3226			Fe (%)			16.46			1.48			16.46			1.48			0			0			0
3227			Ag (g/t)			59.16			0.87			59.16			0.87			0			0			0
3227			Au (g/t)			1.74			0.96			1.74			0.96			0			0			0
3227			Pb (%)			0.62			1.35			0.62			1.35			0			0			0
3227			Zn (%)			1.72			1.05			1.72			1.05			0			0			0
3227			Cu (%)			0.09			0.98			0.09			0.97			0			0.01			-0.01
3227			Fe (%)			21.58			0.54			21.58			0.54			0			0			0
3230			Ag (g/t)			114.03			1.07			63.05			0.8			-50.98			-0.45			-0.28
3230			Au (g/t)			0.79			0.73			0.79			0.73			0			0			0
3230			Pb (%)			1.33			0.24			1.33			0.24			0			0			0
3230			Zn (%)			1.84			0.39			1.84			0.39			0			0			0
3230			Cu (%)			0.05			0.15			0.05			0.13			0			0.03			-0.03
3230			Fe (%)			17.46			0.18			17.46			0.18			0			0			0
3291			Ag (g/t)			104.12			1.22			103.52			1.21			-0.6			-0.01			-0.01

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        443

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3291			Au (g/t)			2.49			1.63			2.28			1.23			-0.21			-0.09			-0.39
3291			Pb (%)			1.55			1.61			1.47			1.46			-0.08			-0.05			-0.15
3291			Zn (%)			2.9			1.55			2.68			1.29			-0.22			-0.07			-0.26
3291			Cu (%)			0.13			1.48			0.13			1.24			-0.01			-0.05			-0.25
3291			Fe (%)			43.74			0.86			42.7			0.81			-1.04			-0.02			-0.04
3292			Ag (g/t)			56.92			0.95			56.92			0.95			0			0			0
3292			Au (g/t)			1.86			1.1			1.75			0.98			-0.11			-0.06			-0.13
3292			Pb (%)			0.81			1.32			0.72			1.1			-0.09			-0.11			-0.22
3292			Zn (%)			2.02			1.04			2.02			1.04			0			0			0
3292			Cu (%)			0.09			1.01			0.08			0.85			-0.01			-0.06			-0.15
3292			Fe (%)			36.06			0.78			36.06			0.78			0			0			0
3293			Ag (g/t)			31.54			1.05			31.54			1.05			0			0			0
3293			Au (g/t)			0.71			1.11			0.71			1.11			0			0			0
3293			Pb (%)			0.75			1.21			0.75			1.21			0			0			0
3293			Zn (%)			1.46			0.95			1.46			0.95			0			0			0
3293			Cu (%)			0.05			1.25			0.05			1.25			0			0.01			-0.01
3293			Fe (%)			21.76			0.63			21.76			0.63			0			0			0
3294			Ag (g/t)			90.46			0.94			90.46			0.94			0			0			0
3294			Au (g/t)			3.81			1.21			3.81			1.21			0			0			0
3294			Pb (%)			0.11			0.65			0.11			0.65			0			0			0
3294			Zn (%)			0.49			1.24			0.49			1.24			0			0			0
3294			Cu (%)			0.13			0.5			0.13			0.5			0			0			0
3294			Fe (%)			80.77			0.88			48.25			0.62			-32.51			-0.4			-0.26
3295			Ag (g/t)			38.4			1.29			38.4			1.29			0			0			0
3295			Au (g/t)			0.38			1.09			0.38			1.09			0			0			0
3295			Pb (%)			0.78			1.35			0.78			1.35			0			0			0
3295			Zn (%)			1.47			1.19			1.47			1.19			0			0			0
3295			Cu (%)			0.1			1.77			0.1			1.77			0			0			0
3295			Fe (%)			23.8			1.69			22.55			1.49			-1.25			-0.05			-0.2
3301			Ag (g/t)			234.86			4.14			130.84			2.14			- 104.02			-0.44			-2
3301			Au (g/t)			1.3			1.49			1.2			1.36			-0.1			-0.08			-0.12
3301			Pb (%)			0.95			1.35			0.92			1.28			-0.03			-0.03			-0.07
3301			Zn (%)			1.7			1.26			1.7			1.26			0			0			0
3301			Cu (%)			0.2			1.96			0.19			1.78			-0.01			-0.06			-0.18
3301			Fe (%)			28.04			0.96			26.62			0.84			-1.42			-0.05			-0.12
3402			Ag (g/t)			13.83			0.35			13.83			0.35			0			0			0
3402			Au (g/t)			0.41			0.89			0.41			0.89			0			0			0
3402			Pb (%)			0.24			0.98			0.24			0.98			0			0			0
3402			Zn (%)			0.7			0.8			0.7			0.8			0			0			0
3402			Cu (%)			0.03			0.44			0.03			0.43			0			0.01			-0.01
3402			Fe (%)			16.67			0.63			16.67			0.63			0			0			0
3409			Ag (g/t)			22.22			0.91			22.22			0.91			0			0			0
3409			Au (g/t)			1.03			0.77			1.03			0.77			0			0			0
3409			Pb (%)			0.42			1.25			0.42			1.25			0			0			0
3409			Zn (%)			1.4			0.99			1.4			0.99			0			0			0
3409			Cu (%)			0.02			1.17			0.02			1.19			0			0.01			0.02
3409			Fe (%)			11.68			0.76			11.68			0.76			0			0			0
3410			Ag (g/t)			11.81			0.98			11.81			0.98			0			0			0
3410			Au (g/t)			0.64			1.49			0.64			1.49			0			0			0
3410			Pb (%)			0.1			0.71			0.1			0.71			0			0			0
3410			Zn (%)			0.73			0.96			0.73			0.96			0			0			0

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        444

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3410			Cu (%)			0.02			0.56			0.02			0.56			0			0.01			0.01
3410			Fe (%)			14.16			0.93			14.16			0.93			0			0			0
3411			Ag (g/t)			28.72			0.73			28.72			0.73			0			0			0
3411			Au (g/t)			1.24			0.87			1.2			0.82			-0.04			-0.03			-0.05
3411			Pb (%)			1			0.82			1			0.82			0			0			0
3411			Zn (%)			2.46			0.9			2.46			0.9			0			0			0
3411			Cu (%)			0.05			0.97			0.05			0.96			0			0.01			-0.01
3411			Fe (%)			22.58			0.73			22.58			0.73			0			0			0
3508			Ag (g/t)			22.47			0.77			22.47			0.77			0			0			0
3508			Au (g/t)			0.87			0.69			0.87			0.69			0			0			0
3508			Pb (%)			0.63			0.97			0.63			0.97			0			0			0
3508			Zn (%)			1.68			0.84			1.63			0.8			-0.05			-0.03			-0.04
3508			Cu (%)			0.02			0.47			0.02			0.48			0			0.02			0.01
3508			Fe (%)			16.79			0.5			16.79			0.5			0			0			0
3600			Ag (g/t)			30.37			0.72			29.73			0.72			-0.64			-0.02			0
3600			Au (g/t)			0.72			0.97			0.72			0.98			0			0			0
3600			Pb (%)			0.88			0.95			0.88			0.95			0			0			0
3600			Zn (%)			2.59			0.96			2.59			0.96			0			0			0
3600			Cu (%)			0.04			1.48			0.04			1.47			0			0.01			-0.02
3600			Fe (%)			16.08			0.84			16.08			0.84			0			0			0
4001			Ag (g/t)			924.75			1.19			722.7			1.12			- 202.05			-0.22			-0.06
4001			Au (g/t)			0.07			0.59			0.07			0.59			0			0			0
4001			Pb (%)			0.21			0.78			0.21			0.78			0			0			0
4001			Zn (%)			0.16			0.57			0.16			0.57			0			0			0
4001			Cu (%)			2.92			0.22			2.92			0.22			0			0			0
4001			Fe (%)			6.94			0.44			6.94			0.44			0			0			0
4002			Ag (g/t)			39.79			1			39.79			1			0			0			0
4002			Au (g/t)			0.01			0.48			0.01			0.48			0			0			0
4002			Pb (%)			2.38			1.04			2.38			1.04			0			0			0
4002			Zn (%)			3.96			0.7			3.96			0.7			0			0			0
4002			Cu (%)			0.09			0.41			0.09			0.41			0			0			0
4002			Fe (%)			2.35			0.37			2.35			0.37			0			0			0
4003			Ag (g/t)			94.14			0.34			94.14			0.34			0			0			0
4003			Au (g/t)			0.07			0.06			0.07			0.06			0			0			0
4003			Pb (%)			4.52			0.26			4.52			0.26			0			0			0
4003			Zn (%)			23.75			0.83			23.75			0.83			0			0			0
4003			Cu (%)			0.3			0.03			0.3			0.03			0			0			0
4003			Fe (%)			3.84			0.7			3.84			0.7			0			0			0

Domain			Element			Uncapped comps mean			Uncapped comps COV			Capped comps mean			Capped comps COV			St difference (mean)			% Difference (mean)			Straight difference (COV)
1003			Ag (g/t)			44.31			2.25			44.31			2.25			0			0			0
1003			Au (g/t)			0.56			1.25			0.56			1.25			0			0			0
1003			Pb (%)			0.42			1.69			0.42			1.69			0			0			0
1003			Zn (%)			1.87			2.28			1.87			2.28			0			0			0
1003			Cu (%)			0.09			1.5			0.09			1.49			0			0			-0.01
1003			Fe (%)			17.31			1.74			17.31			1.74			0			0			0
1004			Ag (g/t)			65.42			0.93			65.42			0.93			0			0			0
1004			Au (g/t)			0.35			0.92			0.35			0.92			0			0			0
1004			Pb (%)			1.79			1.27			1.79			1.27			0			0			0
1004			Zn (%)			2.57			0.71			2.57			0.71			0			0			0

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        445

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1004			Cu (%)			0.06			1.4			0.06			1.4			0			0			0
1004			Fe (%)			14.52			0.87			13.01			0.76			-1.51			-0.1			-0.1
1005			Ag (g/t)			73.82			0.92			73.82			0.92			0			0			0
1005			Au (g/t)			0.47			0.53			0.47			0.53			0			0			0
1005			Pb (%)			0.27			0.37			0.27			0.37			0			0			0
1005			Zn (%)			1.05			0.37			1.05			0.37			0			0			0
1005			Cu (%)			0.03			0.9			0.03			0.9			0			0			0
1005			Fe (%)			10.88			0.77			10.88			0.77			0			0			0
1006			Ag (g/t)			30.75			0.65			30.75			0.65			0			0			0
1006			Au (g/t)			0.38			0.8			0.38			0.8			0			0			0
1006			Pb (%)			0.45			0.74			0.45			0.74			0			0			0
1006			Zn (%)			1.45			0.75			1.45			0.75			0			0			0
1006			Cu (%)			0.02			0.71			0.02			0.71			0			0			0
1006			Fe (%)			11.99			0.97			11.99			0.97			0			0			0
1007			Ag (g/t)			200.32			1.61			177.74			1.57			-22.58			-0.11			-0.05
1007			Au (g/t)			0.64			0.58			0.64			0.58			0			0			0
1007			Pb (%)			0.53			0.94			0.53			0.94			0			0			0
1007			Zn (%)			2.51			1.37			2.51			1.37			0			0			0
1007			Cu (%)			0.23			1.41			0.23			1.41			0			0			0
1007			Fe (%)			14.65			0.63			13.61			0.55			-1.04			-0.07			-0.08
1011			Ag (g/t)			57.16			0.72			57.16			0.72			0			0			0
1011			Au (g/t)			1.59			0.95			1.59			0.95			0			0			0
1011			Pb (%)			0.47			0.69			0.47			0.69			0			0			0
1011			Zn (%)			0.92			0.73			0.92			0.73			0			0			0
1011			Cu (%)			0.03			0.89			0.03			0.89			0			0			0
1011			Fe (%)			35.59			0.87			35.59			0.87			0			0			0
1012			Ag (g/t)			129.23			0.38			129.23			0.38			0			0			0
1012			Au (g/t)			1			0.86			1			0.86			0			0			0
1012			Pb (%)			1.24			0.86			1.24			0.86			0			0			0
1012			Zn (%)			2.26			0.93			2.26			0.93			0			0			0
1012			Cu (%)			0.31			0.63			0.31			0.63			0			0			0
1012			Fe (%)			22.78			0.46			22.78			0.46			0			0			0
1013			Ag (g/t)			29.75			0.7			29.75			0.7			0			0			0
1013			Au (g/t)			0.12			0.91			0.12			0.91			0			0			0
1013			Pb (%)			0.49			1.34			0.49			1.34			0			0			0
1013			Zn (%)			1.07			1.39			1.07			1.39			0			0			0
1013			Cu (%)			0.03			0.38			0.03			0.38			0			0			0
1013			Fe (%)			17.39			0.76			17.39			0.76			0			0			0
1030			Ag (g/t)			130.13			1.25			130.13			1.25			0			0			0
1030			Au (g/t)			2.29			1.92			2.29			1.92			0			0			0
1030			Pb (%)			1.64			1.2			1.64			1.2			0			0			0
1030			Zn (%)			7.65			1.37			7.65			1.37			0			0			0
1030			Cu (%)			0.21			1.85			0.21			1.85			0			0			0
1030			Fe (%)			21.76			1.09			21.76			1.09			0			0			0
1031			Ag (g/t)			99.97			1.34			96.85			1.32			-3.13			-0.03			-0.02
1031			Au (g/t)			1.31			3.26			1.31			3.26			0			0			0
1031			Pb (%)			0.88			1.36			0.88			1.36			0			0			0
1031			Zn (%)			2.43			2.11			2.43			2.11			0			0			0
1031			Cu (%)			0.06			2.25			0.06			2.25			0			0.01			0
1031			Fe (%)			19.63			1.85			19.63			1.85			0			0			0
1033			Ag (g/t)			20.62			0.83			20.62			0.83			0			0			0

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        446

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1033			Au (g/t)			0.99			0.96			0.99			0.96			0			0			0
1033			Pb (%)			0.24			0.71			0.24			0.71			0			0			0
1033			Zn (%)			0.36			1.06			0.36			1.06			0			0			0
1033			Cu (%)			0.03			1.57			0.03			1.54			0			0.02			-0.03
1033			Fe (%)			12.42			0.7			12.42			0.7			0			0			0
1040			Ag (g/t)			85.8			1.05			85.8			1.05			0			0			0
1040			Au (g/t)			2.58			1.34			2.45			1.34			-0.12			-0.05			0
1040			Pb (%)			2.14			1.83			2.14			1.83			0			0			0
1040			Zn (%)			3.84			2.07			3.84			2.07			0			0			0
1040			Cu (%)			0.06			1.07			0.06			1.06			0			0.01			0
1040			Fe (%)			24.28			0.81			24.28			0.81			0			0			0
1041			Ag (g/t)			148.18			1.49			148.18			1.49			0			0			0
1041			Au (g/t)			1.74			2.08			1.67			2.12			-0.08			-0.04			0.05
1041			Pb (%)			2.16			1.49			2.16			1.49			0			0			0
1041			Zn (%)			3.57			1.43			3.49			1.43			-0.08			-0.02			0
1041			Cu (%)			0.04			1.26			0.04			1.26			0			0			0
1041			Fe (%)			19.96			0.96			19.96			0.96			0			0			0
1050			Ag (g/t)			46.76			1.27			46.76			1.27			0			0			0
1050			Au (g/t)			2.65			1.5			2.65			1.5			0			0			0
1050			Pb (%)			0.93			2.06			0.93			2.06			0			0			0
1050			Zn (%)			2.91			2.03			2.75			1.97			-0.17			-0.06			-0.06
1050			Cu (%)			0.35			1.48			0.35			1.48			0			0			0
1050			Fe (%)			22.58			1.29			22.58			1.29			0			0			0
1051			Ag (g/t)			27.98			0.81			27.98			0.81			0			0			0
1051			Au (g/t)			0.3			0.46			0.3			0.46			0			0			0
1051			Pb (%)			0.11			0.63			0.11			0.63			0			0			0
1051			Zn (%)			0.04			0.91			0.04			0.91			0			0			0
1051			Cu (%)			0.01			0.84			0.01			0.73			0			0.06			-0.11
1051			Fe (%)			3.67			0.1			3.67			0.1			0			0			0
1052			Ag (g/t)			48.42			1.37			48.42			1.37			0			0			0
1052			Au (g/t)			0.78			1.32			0.78			1.32			0			0			0
1052			Pb (%)			0.82			1.23			0.82			1.23			0			0			0
1052			Zn (%)			0.87			1.27			0.87			1.27			0			0			0
1052			Cu (%)			0.19			2.84			0.11			1.39			-0.08			-0.41			-1.45
1052			Fe (%)			17.67			0.98			17.67			0.98			0			0			0
1053			Ag (g/t)			39.73			0.57			39.73			0.57			0			0			0
1053			Au (g/t)			0.84			0.3			0.84			0.3			0			0			0
1053			Pb (%)			0.68			1.04			0.68			1.04			0			0			0
1053			Zn (%)			1.92			1.3			1.92			1.3			0			0			0
1053			Cu (%)			0.07			0.51			0.07			0.51			0			0			0
1053			Fe (%)			13.65			0.3			13.65			0.3			0			0			0
1054			Ag (g/t)			22.97			1.34			22.97			1.34			0			0			0
1054			Au (g/t)			0.05			1.19			0.05			1.19			0			0			0
1054			Pb (%)			0.97			1.7			0.97			1.7			0			0			0
1054			Zn (%)			1.98			2.14			1.98			2.14			0			0			0
1054			Cu (%)			0.25			0.86			0.25			0.86			0			0			0
1054			Fe (%)			9.14			1.48			9.14			1.48			0			0			0
1055			Ag (g/t)			296.3			0.68			296.3			0.68			0			0			0
1055			Au (g/t)			0.72			0.7			0.72			0.7			0			0			0
1055			Pb (%)			4.74			0.7			4.74			0.7			0			0			0
1055			Zn (%)			4.53			0.85			4.53			0.85			0			0			0

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        447

---

1055			Cu (%)			0.07			0.92			0.07			0.92			0			0			0
1055			Fe (%)			20.67			0.71			20.67			0.71			0			0			0
1056			Ag (g/t)			49.3			0.64			49.3			0.64			0			0			0
1056			Au (g/t)			0.71			0.98			0.71			0.98			0			0			0
1056			Pb (%)			0.69			1.57			0.69			1.57			0			0			0
1056			Zn (%)			0.65			1.12			0.65			1.12			0			0			0
1056			Cu (%)			0.46			0.61			0.46			0.61			0			0			0
1056			Fe (%)			40.23			0.91			37.39			0.81			-2.84			-0.07			-0.1
1061			Ag (g/t)			122.19			0.6			122.19			0.6			0			0			0
1061			Au (g/t)			0.62			0.78			0.62			0.78			0			0			0
1061			Pb (%)			2.82			0.59			2.82			0.59			0			0			0
1061			Zn (%)			6.75			0.85			6.75			0.85			0			0			0
1061			Cu (%)			0.03			0.11			0.03			0.03			0			0.14			-0.08
1061			Fe (%)			28.03			0.52			28.03			0.52			0			0			0
1062			Ag (g/t)			22.45			1.35			22.45			1.35			0			0			0
1062			Au (g/t)			0.06			0.85			0.06			0.85			0			0			0
1062			Pb (%)			5.21			1.41			5.21			1.41			0			0			0
1062			Zn (%)			10.88			1.41			9.91			1.41			-0.97			-0.09			0
1062			Cu (%)			0.22			0.66			0.22			0.66			0			0.01			0
1062			Fe (%)			17.55			0.74			17.55			0.74			0			0			0
1063			Ag (g/t)			3.08			0.77			3.08			0.77			0			0			0
1063			Au (g/t)			7.94			0.48			7.94			0.48			0			0			0
1063			Pb (%)			0.02			0.64			0.02			0.64			0			0			0
1063			Zn (%)			0.03			0.86			0.03			0.86			0			0			0
1063			Cu (%)			0.02			0.54			0.02			0.54			0			0			0
1063			Fe (%)			34.44			0.83			34.44			0.83			0			0			0
1090			Ag (g/t)			258.26			1.94			258.26			1.94			0			0			0
1090			Au (g/t)			3.77			1.31			3.77			1.31			0			0			0
1090			Pb (%)			2.87			2.39			2.87			2.39			0			0			0
1090			Zn (%)			9.41			1.89			9.35			1.88			-0.06			-0.01			-0.01
1090			Cu (%)			0.24			1.93			0.24			1.93			0			0			-0.01
1090			Fe (%)			44.69			0.99			43.29			0.89			-1.4			-0.03			-0.1
1091			Ag (g/t)			432.92			1.61			394.35			1.27			-38.57			-0.09			-0.34
1091			Au (g/t)			1.68			1.17			1.65			1.11			-0.04			-0.02			-0.06
1091			Pb (%)			3.07			1.27			2.74			1.04			-0.33			-0.11			-0.23
1091			Zn (%)			4.68			1.04			4.43			0.92			-0.25			-0.05			-0.12
1091			Cu (%)			0.14			2.22			0.14			2.13			0			-0.03			-0.09
1091			Fe (%)			32.27			0.77			32.27			0.77			0			0			0
1092			Ag (g/t)			91.96			1.89			91.96			1.89			0			0			0
1092			Au (g/t)			4.19			1.9			4.19			1.9			0			0			0
1092			Pb (%)			1.25			1.14			1.25			1.14			0			0			0
1092			Zn (%)			3.2			1.75			3.2			1.75			0			0			0
1092			Cu (%)			0.27			1.67			0.27			1.66			0			0			0
1092			Fe (%)			43.15			1.27			43.15			1.27			0			0			0
1093			Ag (g/t)			329.44			2.49			329.44			2.49			0			0			0
1093			Au (g/t)			2.51			1.53			2.51			1.53			0			0			0
1093			Pb (%)			1.04			1.3			1.04			1.3			0			0			0
1093			Zn (%)			3.26			1.66			3.26			1.66			0			0			0
1093			Cu (%)			0.22			1.55			0.22			1.55			0			0			0
1093			Fe (%)			29.68			0.95			29.68			0.95			0			0			0
1094			Ag (g/t)			105.6			1.04			105.6			1.04			0			0			0

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        448

---

1094			Au (g/t)			6.21			1.2			6.21			1.2			0			0			0
1094			Pb (%)			1.19			0.77			1.19			0.77			0			0			0
1094			Zn (%)			1.51			1.11			1.51			1.11			0			0			0
1094			Cu (%)			0.2			1.24			0.2			1.23			0			0			0
1094			Fe (%)			49.41			0.71			49.41			0.71			0			0			0
1201			Ag (g/t)			1095.83			1.65			1058.78			1.64			-37.05			-0.03			-0.01
1201			Au (g/t)			5.78			1.17			5.78			1.17			0			0			0
1201			Pb (%)			10.26			1.72			10.21			1.73			-0.05			-0.01			0.01
1201			Zn (%)			12.92			1.01			12.87			1.01			-0.05			0			0
1201			Cu (%)			0.22			0.93			0.22			0.93			0			0			0
1201			Fe (%)			78.75			0.88			78.75			0.88			0			0			0
1202			Ag (g/t)			110.65			2.3			110.65			2.3			0			0			0
1202			Au (g/t)			1.96			2.99			1.63			2.84			-0.33			-0.17			-0.15
1202			Pb (%)			1.17			2.25			1.17			2.25			0			0			0
1202			Zn (%)			2.74			1.89			2.74			1.89			0			0			0
1202			Cu (%)			0.1			2.71			0.1			2.7			0			0			-0.01
1202			Fe (%)			19.41			1.52			14.19			1.09			-5.21			-0.27			-0.43
1203			Ag (g/t)			71.92			2.04			71.92			2.04			0			0			0
1203			Au (g/t)			1.48			2.78			1.48			2.78			0			0			0
1203			Pb (%)			1.4			1.96			1.4			1.96			0			0			0
1203			Zn (%)			3.14			1.93			3.14			1.93			0			0			0
1203			Cu (%)			0.05			1.59			0.05			1.58			0			0			0
1203			Fe (%)			18.33			1.61			18.33			1.61			0			0			0
1208			Ag (g/t)			21.12			0.65			21.12			0.65			0			0			0
1208			Au (g/t)			0.58			0.32			0.58			0.32			0			0			0
1208			Pb (%)			0.16			0.71			0.16			0.71			0			0			0
1208			Zn (%)			0.16			0.97			0.16			0.97			0			0			0
1208			Cu (%)			0.04			0.89			0.04			0.87			0			0.01			-0.02
1208			Fe (%)			13.18			0.58			13.18			0.58			0			0			0
1209			Ag (g/t)			20.16			0.83			20.16			0.83			0			0			0
1209			Au (g/t)			0.57			1.74			0.57			1.74			0			0			0
1209			Pb (%)			0.36			0.7			0.36			0.7			0			0			0
1209			Zn (%)			1.34			0.82			1.34			0.82			0			0			0
1209			Cu (%)			0.18			1.75			0.18			1.75			0			0			0
1209			Fe (%)			7.77			0.58			7.77			0.58			0			0			0
1304			Ag (g/t)			29.04			0.4			29.04			0.4			0			0			0
1304			Au (g/t)			1.12			0.79			1.12			0.79			0			0			0
1304			Pb (%)			1.41			0.99			1.41			0.99			0			0			0
1304			Zn (%)			1.28			0.84			1.28			0.84			0			0			0
1304			Cu (%)			0.07			0.4			0.07			0.37			0			0.02			-0.02
1304			Fe (%)			24.37			0.63			24.37			0.63			0			0			0
1305			Ag (g/t)			54.67			1.31			54.67			1.31			0			0			0
1305			Au (g/t)			2.3			1.54			2.3			1.54			0			0			0
1305			Pb (%)			0.58			1.46			0.58			1.46			0			0			0
1305			Zn (%)			1.03			1.38			1.03			1.38			0			0			0
1305			Cu (%)			0.16			1.05			0.16			1.05			0			0			0
1305			Fe (%)			31.91			0.82			31.06			0.75			-0.85			-0.03			-0.07
1306			Ag (g/t)			144.14			1.15			144.14			1.15			0			0			0
1306			Au (g/t)			2.74			1.74			2.74			1.74			0			0			0
1306			Pb (%)			1.37			1.13			1.37			1.13			0			0			0
1306			Zn (%)			2.48			1.32			2.48			1.32			0			0			0

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        449

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1306			Cu (%)			0.21			1.45			0.21			1.45			0			0			0
1306			Fe (%)			31.39			1.19			31.39			1.19			0			0			0
1310			Ag (g/t)			46.17			0.9			46.17			0.9			0			0			0
1310			Au (g/t)			2.67			1.27			2.67			1.27			0			0			0
1310			Pb (%)			0.99			1.3			0.99			1.3			0			0			0
1310			Zn (%)			2.32			1.67			2.32			1.67			0			0			0
1310			Cu (%)			0.17			1.16			0.17			1.16			0			0			0
1310			Fe (%)			33.37			0.95			33.37			0.95			0			0			0
1311			Ag (g/t)			43.29			1.11			43.29			1.11			0			0			0
1311			Au (g/t)			1.31			2.01			1.31			2.01			0			0			0
1311			Pb (%)			0.12			0.62			0.12			0.62			0			0			0
1311			Zn (%)			0.43			0.76			0.43			0.76			0			0			0
1311			Cu (%)			0.1			0.96			0.1			0.96			0			0			0
1311			Fe (%)			32.96			1.12			27.39			0.91			-5.57			-0.17			-0.21
1312			Ag (g/t)			44.6			1.78			44.6			1.78			0			0			0
1312			Au (g/t)			0.83			1.78			0.83			1.78			0			0			0
1312			Pb (%)			0.58			0.96			0.58			0.96			0			0			0
1312			Zn (%)			3.18			2.43			3.18			2.43			0			0			0
1312			Cu (%)			0.1			1.61			0.1			1.58			0			0.01			-0.03
1312			Fe (%)			15.19			1.13			15.19			1.13			0			0			0
1313			Ag (g/t)			38.1			1.89			38.1			1.89			0			0			0
1313			Au (g/t)			0.21			1.02			0.21			1.02			0			0			0
1313			Pb (%)			2.75			2.07			2.75			2.07			0			0			0
1313			Zn (%)			2.44			0.86			2.44			0.86			0			0			0
1313			Cu (%)			0.04			0.87			0.04			0.89			0			0.04			0.01
1313			Fe (%)			10.56			0.5			10.56			0.5			0			0			0
1314			Ag (g/t)			24.01			0.88			24.01			0.88			0			0			0
1314			Au (g/t)			1.25			0.92			1.25			0.92			0			0			0
1314			Pb (%)			0.32			1.22			0.32			1.22			0			0			0
1314			Zn (%)			0.49			1.55			0.49			1.55			0			0			0
1314			Cu (%)			0.12			0.91			0.12			0.9			0			0			-0.01
1314			Fe (%)			23.93			0.5			23.93			0.5			0			0			0
1331			Ag (g/t)			39.9			0.84			39.9			0.84			0			0			0
1331			Au (g/t)			1.49			1.84			1.49			1.84			0			0			0
1331			Pb (%)			2.13			1.74			2.13			1.74			0			0			0
1331			Zn (%)			2.18			0.93			2.18			0.93			0			0			0
1331			Cu (%)			0.11			1.43			0.11			1.43			0			0.01			-0.01
1331			Fe (%)			21.34			0.84			21.34			0.84			0			0			0
1341			Ag (g/t)			41.69			1.47			41.69			1.47			0			0			0
1341			Au (g/t)			0.75			1.52			0.75			1.52			0			0			0
1341			Pb (%)			2.68			2.45			2.68			2.45			0			0			0
1341			Zn (%)			4.24			1.69			4.24			1.69			0			0			0
1341			Cu (%)			0.12			1.8			0.12			1.79			0			0.01			-0.01
1341			Fe (%)			18.34			0.81			18.34			0.81			0			0			0
1403			Ag (g/t)			36.95			0.81			36.95			0.81			0			0			0
1403			Au (g/t)			0.89			1.45			0.89			1.45			0			0			0
1403			Pb (%)			3.68			1.42			3.68			1.42			0			0			0
1403			Zn (%)			3.36			0.74			3.36			0.74			0			0			0
1403			Cu (%)			0.09			0.65			0.09			0.65			0			0.01			0
1403			Fe (%)			32.88			0.55			32.88			0.55			0			0			0
1404			Ag (g/t)			22.84			1.17			22.84			1.17			0			0			0

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        450

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1404			Au (g/t)			0.21			1.37			0.21			1.37			0			0			0
1404			Pb (%)			0.39			0.83			0.39			0.83			0			0			0
1404			Zn (%)			1.39			1.27			1.39			1.27			0			0			0
1404			Cu (%)			0.06			2.39			0.07			2.36			0			0.01			-0.02
1404			Fe (%)			11.09			0.58			11.09			0.58			0			0			0
1407			Ag (g/t)			45			1.25			45			1.25			0			0			0
1407			Au (g/t)			2.57			2.62			1.99			2.26			-0.58			-0.22			-0.36
1407			Pb (%)			0.98			1.3			0.98			1.3			0			0			0
1407			Zn (%)			1.76			1.18			1.76			1.18			0			0			0
1407			Cu (%)			0.15			1.55			0.15			1.54			0			0.01			-0.01
1407			Fe (%)			39.15			1.21			39.15			1.21			0			0			0
1501			Ag (g/t)			207.96			1.23			207.96			1.23			0			0			0
1501			Au (g/t)			2.23			1.2			2.14			1.14			-0.09			-0.04			-0.06
1501			Pb (%)			2.64			1.27			2.64			1.27			0			0			0
1501			Zn (%)			5.15			1.32			5.15			1.32			0			0			0
1501			Cu (%)			0.08			1.19			0.08			1.18			0			0.01			-0.01
1501			Fe (%)			37.53			0.89			37.53			0.89			0			0			0
1502			Ag (g/t)			275.26			1.53			275.26			1.53			0			0			0
1502			Au (g/t)			1.2			1.02			1.2			1.02			0			0			0
1502			Pb (%)			2.51			1.26			2.51			1.26			0			0			0
1502			Zn (%)			6.3			1.23			6.19			1.24			-0.12			-0.02			0.01
1502			Cu (%)			0.06			1.49			0.06			1.48			0			0.01			-0.01
1502			Fe (%)			24.74			0.86			24.74			0.86			0			0			0
1503			Ag (g/t)			236.47			1.4			230.2			1.36			-6.27			-0.03			-0.04
1503			Au (g/t)			1.2			1.04			1.2			1.04			0			0			0
1503			Pb (%)			2.48			1.52			2.38			1.44			-0.1			-0.04			-0.07
1503			Zn (%)			4.92			1.32			4.92			1.32			0			0			0
1503			Cu (%)			0.13			1.64			0.13			1.63			0			0			-0.01
1503			Fe (%)			32.7			0.96			32.7			0.96			0			0			0
1504			Ag (g/t)			252.69			1.33			252.69			1.33			0			0			0
1504			Au (g/t)			2.36			1.26			1.9			1.02			-0.46			-0.2			-0.25
1504			Pb (%)			2.32			1.52			2.32			1.52			0			0			0
1504			Zn (%)			3.79			1.46			3.79			1.46			0			0			0
1504			Cu (%)			0.05			0.92			0.05			0.92			0			0			0
1504			Fe (%)			30.13			0.86			30.13			0.86			0			0			0
1505			Ag (g/t)			102.06			1.06			102.06			1.06			0			0			0
1505			Au (g/t)			0.75			1.26			0.75			1.26			0			0			0
1505			Pb (%)			0.95			1.44			0.95			1.44			0			0			0
1505			Zn (%)			1.73			1.26			1.73			1.26			0			0			0
1505			Cu (%)			0.04			1.28			0.04			1.24			0			0.02			-0.04
1505			Fe (%)			19.15			1.13			19.15			1.13			0			0			0
1601			Ag (g/t)			21.82			1.09			21.82			1.09			0			0			0
1601			Au (g/t)			0.16			0.71			0.16			0.71			0			0			0
1601			Pb (%)			1.28			1.46			1.28			1.46			0			0			0
1601			Zn (%)			1.81			0.76			1.81			0.76			0			0			0
1601			Cu (%)			0.09			0.98			0.09			0.98			0			0			0
1601			Fe (%)			13.68			0.55			13.68			0.55			0			0			0
1614			Ag (g/t)			14.43			0.88			14.43			0.88			0			0			0
1614			Au (g/t)			0.47			0.91			0.47			0.91			0			0			0
1614			Pb (%)			0.1			0.78			0.1			0.78			0			0			0
1614			Zn (%)			0.09			0.69			0.09			0.69			0			0			0

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        451

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1614			Cu (%)			0.04			0.99			0.04			0.98			0			0			0
1614			Fe (%)			12.88			0.76			12.88			0.76			0			0			0
1631			Ag (g/t)			87.23			0.81			87.23			0.81			0			0			0
1631			Au (g/t)			1.74			0.98			1.74			0.98			0			0			0
1631			Pb (%)			1.67			0.8			1.67			0.8			0			0			0
1631			Zn (%)			2.16			1.17			2.16			1.17			0			0			0
1631			Cu (%)			0.06			0.77			0.06			0.75			0			0.02			-0.02
1631			Fe (%)			35.63			0.68			35.63			0.68			0			0			0
1632			Ag (g/t)			77.03			1.55			64.12			1.28			-12.91			-0.17			-0.27
1632			Au (g/t)			3.18			1.39			2.81			1.25			-0.38			-0.12			-0.14
1632			Pb (%)			1.44			1.94			0.82			0.77			-0.62			-0.43			-1.18
1632			Zn (%)			1.97			0.67			1.97			0.67			0			0			0
1632			Cu (%)			0.13			1.4			0.12			1.22			-0.02			-0.11			-0.18
1632			Fe (%)			33.16			0.99			33.16			0.99			0			0			0
2002			Ag (g/t)			72.51			1.87			72.51			1.87			0			0			0
2002			Au (g/t)			3.65			1.97			3.65			1.97			0			0			0
2002			Pb (%)			2.27			2.03			2.27			2.03			0			0			0
2002			Zn (%)			3.34			2.05			3.34			2.05			0			0			0
2002			Cu (%)			0.08			1.35			0.08			1.34			0			0.01			-0.02
2002			Fe (%)			31.48			1.01			31.48			1.01			0			0			0
2004			Ag (g/t)			186.38			1.23			186.38			1.23			0			0			0
2004			Au (g/t)			3.2			1.02			3.2			1.02			0			0			0
2004			Pb (%)			1.23			1.32			1.23			1.32			0			0			0
2004			Zn (%)			4.94			1.02			4.94			1.02			0			0			0
2004			Cu (%)			0.3			1.28			0.31			1.27			0			0			-0.01
2004			Fe (%)			43.24			0.72			43.24			0.72			0			0			0
2005			Ag (g/t)			19.57			0.85			19.57			0.85			0			0			0
2005			Au (g/t)			0.26			1.68			0.26			1.68			0			0			0
2005			Pb (%)			0.41			1.48			0.41			1.48			0			0			0
2005			Zn (%)			1.62			1.18			1.62			1.18			0			0			0
2005			Cu (%)			0.01			0.63			0.01			0.62			0			0.03			-0.01
2005			Fe (%)			13.43			0.58			13.43			0.58			0			0			0
2006			Ag (g/t)			32.99			0.89			32.99			0.89			0			0			0
2006			Au (g/t)			0.62			2.32			0.62			2.32			0			0			0
2006			Pb (%)			2.81			1.9			1.86			1.52			-0.95			-0.34			-0.39
2006			Zn (%)			1.74			1.09			1.74			1.09			0			0			0
2006			Cu (%)			0.03			1.3			0.03			1.28			0			0.01			-0.01
2006			Fe (%)			12.1			1.05			12.1			1.05			0			0			0
2009			Ag (g/t)			69.51			0.79			69.51			0.79			0			0			0
2009			Au (g/t)			2.71			1.04			2.71			1.04			0			0			0
2009			Pb (%)			1.69			1.41			1.69			1.41			0			0			0
2009			Zn (%)			2.63			0.63			2.63			0.63			0			0			0
2009			Cu (%)			0.1			1.1			0.1			1.09			0			0.01			-0.01
2009			Fe (%)			42.65			0.8			42.65			0.8			0			0			0
2012			Ag (g/t)			74.49			1.04			71.77			1.04			-2.73			-0.04			0
2012			Au (g/t)			1.43			1.43			1.43			1.43			0			0			0
2012			Pb (%)			2.26			1.41			2.26			1.41			0			0			0
2012			Zn (%)			6.95			1.41			6.95			1.41			0			0			0
2012			Cu (%)			0.08			1.91			0.08			1.89			0			0.01			-0.01
2012			Fe (%)			25.01			0.84			25.01			0.84			0			0			0
2018			Ag (g/t)			176.5			1.01			176.5			1.01			0			0			0

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        452

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2018			Au (g/t)			5.06			1.19			5.06			1.19			0			0			0
2018			Pb (%)			1.69			0.76			1.69			0.76			0			0			0
2018			Zn (%)			9.29			0.8			8.88			0.8			-0.41			-0.04			0
2018			Cu (%)			0.1			1.29			0.1			1.28			0			0.01			-0.01
2018			Fe (%)			53.38			1.03			53.38			1.03			0			0			0
2019			Ag (g/t)			159.32			0.73			159.32			0.73			0			0			0
2019			Au (g/t)			4.63			1.02			4.63			1.02			0			0			0
2019			Pb (%)			0.51			0.6			0.51			0.6			0			0			0
2019			Zn (%)			1.43			0.98			1.43			0.98			0			0			0
2019			Cu (%)			0.35			0.7			0.35			0.7			0			0			0
2019			Fe (%)			50.84			1.05			50.84			1.05			0			0			0
2020			Ag (g/t)			123.27			1.16			123.27			1.16			0			0			0
2020			Au (g/t)			13.42			1.38			10.7			1.26			-2.72			-0.2			-0.12
2020			Pb (%)			1.34			1.94			1.34			1.94			0			0			0
2020			Zn (%)			2.78			1.77			2.78			1.77			0			0			0
2020			Cu (%)			0.19			0.91			0.19			0.9			0			0			0
2020			Fe (%)			53.1			0.8			53.1			0.8			0			0			0
2026			Ag (g/t)			18.6			0.86			18.6			0.86			0			0			0
2026			Au (g/t)			0.77			0.99			0.77			0.99			0			0			0
2026			Pb (%)			0.89			1.5			0.89			1.5			0			0			0
2026			Zn (%)			2.17			1.24			2.17			1.24			0			0			0
2026			Cu (%)			0.02			1.17			0.02			0.97			0			0.2			-0.2
2026			Fe (%)			16.5			0.69			16.5			0.69			0			0			0
2091			Ag (g/t)			246.09			1.45			221.43			1.22			-24.66			-0.1			-0.23
2091			Au (g/t)			11.33			1.93			10.15			1.5			-1.18			-0.1			-0.42
2091			Pb (%)			2.03			1.21			2			1.17			-0.03			-0.01			-0.05
2091			Zn (%)			7.39			1.34			7.11			1.21			-0.28			-0.04			-0.14
2091			Cu (%)			0.39			1.71			0.37			1.51			-0.02			-0.04			-0.2
2091			Fe (%)			74.85			0.89			73.71			0.86			-1.13			-0.02			-0.03
2101			Ag (g/t)			40.64			1.25			40.64			1.25			0			0			0
2101			Au (g/t)			1.49			1.77			1.49			1.77			0			0			0
2101			Pb (%)			0.41			0.79			0.41			0.79			0			0			0
2101			Zn (%)			2.3			1.18			2.13			1.22			-0.16			-0.07			0.05
2101			Cu (%)			0.05			1.76			0.05			1.75			0			0.01			-0.02
2101			Fe (%)			12.82			1.12			12.82			1.12			0			0			0
2103			Ag (g/t)			69.05			1.46			69.05			1.46			0			0			0
2103			Au (g/t)			2.47			2.25			2.47			2.25			0			0			0
2103			Pb (%)			1.48			1.72			1.48			1.72			0			0			0
2103			Zn (%)			5.09			1.7			5.09			1.7			0			0			0
2103			Cu (%)			0.1			1.75			0.11			1.74			0			0			-0.01
2103			Fe (%)			30.74			1.48			30.74			1.48			0			0			0
2104			Ag (g/t)			99.57			1.71			98.01			1.68			-1.56			-0.02			-0.03
2104			Au (g/t)			3.76			1.64			3.52			1.51			-0.24			-0.06			-0.13
2104			Pb (%)			1.36			1.32			1.32			1.24			-0.05			-0.03			-0.08
2104			Zn (%)			4.24			1.28			4.08			1.21			-0.16			-0.04			-0.08
2104			Cu (%)			0.18			1.63			0.18			1.54			-0.01			-0.03			-0.08
2104			Fe (%)			37.59			1.21			37.59			1.21			0			0			0
2106			Ag (g/t)			47.41			0.87			47.41			0.87			0			0			0
2106			Au (g/t)			2.82			1.35			2.82			1.35			0			0			0
2106			Pb (%)			0.94			1			0.94			1			0			0			0
2106			Zn (%)			2.91			1.18			2.91			1.18			0			0			0

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        453

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2106			Cu (%)			0.2			2.4			0.2			2.39			0			0			0
2106			Fe (%)			25.86			0.68			25.86			0.68			0			0			0
2113			Ag (g/t)			37.39			0.94			37.39			0.94			0			0			0
2113			Au (g/t)			0.94			1.03			0.94			1.03			0			0			0
2113			Pb (%)			0.92			0.7			0.92			0.7			0			0			0
2113			Zn (%)			1.76			0.92			1.76			0.92			0			0			0
2113			Cu (%)			0.08			0.68			0.08			0.68			0			0			0
2113			Fe (%)			17.82			0.87			17.82			0.87			0			0			0
2115			Ag (g/t)			25.11			0.62			25.11			0.62			0			0			0
2115			Au (g/t)			0.37			0.53			0.37			0.53			0			0			0
2115			Pb (%)			0.85			1.14			0.85			1.14			0			0			0
2115			Zn (%)			1.92			1.08			1.92			1.08			0			0			0
2115			Cu (%)			0.05			0.67			0.05			0.66			0			0.01			-0.01
2115			Fe (%)			16.47			0.86			16.47			0.86			0			0			0
2117			Ag (g/t)			140.7			1.26			140.7			1.26			0			0			0
2117			Au (g/t)			4.72			1.66			4.72			1.66			0			0			0
2117			Pb (%)			1.21			1.26			1.21			1.26			0			0			0
2117			Zn (%)			1.77			1.07			1.77			1.07			0			0			0
2117			Cu (%)			0.26			1.69			0.26			1.69			0			0			-0.01
2117			Fe (%)			34.61			1.01			34.61			1.01			0			0			0
2118			Ag (g/t)			58.32			0.72			58.32			0.72			0			0			0
2118			Au (g/t)			2.36			1.02			2.36			1.02			0			0			0
2118			Pb (%)			1.22			0.69			1.22			0.69			0			0			0
2118			Zn (%)			3.19			0.62			3.19			0.62			0			0			0
2118			Cu (%)			0.15			1.43			0.15			1.42			0			0			-0.01
2118			Fe (%)			38.73			1.04			38.73			1.04			0			0			0
2120			Ag (g/t)			169.33			0.54			169.33			0.54			0			0			0
2120			Au (g/t)			5.39			0.7			5.39			0.7			0			0			0
2120			Pb (%)			1.01			0.83			1.01			0.83			0			0			0
2120			Zn (%)			3.01			1.08			3.01			1.08			0			0			0
2120			Cu (%)			0.37			0.68			0.37			0.68			0			0			0
2120			Fe (%)			41.91			0.61			41.91			0.61			0			0			0
2122			Ag (g/t)			236.89			1.81			236.89			1.81			0			0			0
2122			Au (g/t)			1.65			1.28			1.65			1.28			0			0			0
2122			Pb (%)			1.31			1.48			1.31			1.48			0			0			0
2122			Zn (%)			0.68			1.2			0.68			1.2			0			0			0
2122			Cu (%)			0.08			1			0.08			0.98			0			0.01			-0.02
2122			Fe (%)			24.71			0.93			24.71			0.93			0			0			0
2127			Ag (g/t)			19.8			0.77			19.8			0.77			0			0			0
2127			Au (g/t)			0.07			0.83			0.07			0.83			0			0			0
2127			Pb (%)			0.12			0.69			0.12			0.69			0			0			0
2127			Zn (%)			1.27			0.95			1.27			0.95			0			0			0
2127			Cu (%)			0.1			0.89			0.1			0.89			0			0			0
2127			Fe (%)			9.84			0.3			9.84			0.3			0			0			0
2130			Ag (g/t)			19.55			0.64			19.55			0.64			0			0			0
2130			Au (g/t)			0.9			1.49			0.9			1.49			0			0			0
2130			Pb (%)			1.43			0.97			1.43			0.97			0			0			0
2130			Zn (%)			2.07			1.45			2.07			1.45			0			0			0
2130			Cu (%)			0.15			1.54			0.15			1.54			0			0			0
2130			Fe (%)			11.62			0.55			11.62			0.55			0			0			0
2131			Ag (g/t)			21.9			0.74			21.9			0.74			0			0			0

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        454

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2131			Au (g/t)			1.28			0.9			1.28			0.9			0			0			0
2131			Pb (%)			0.69			1.14			0.69			1.14			0			0			0
2131			Zn (%)			1.53			1.34			1.53			1.34			0			0			0
2131			Cu (%)			0.03			0.74			0.03			0.74			0			0.01			0
2131			Fe (%)			23.49			0.61			23.49			0.61			0			0			0
2132			Ag (g/t)			22.7			1.36			22.7			1.36			0			0			0
2132			Au (g/t)			0.47			1.82			0.47			1.82			0			0			0
2132			Pb (%)			0.64			2.15			0.64			2.15			0			0			0
2132			Zn (%)			1.14			1.09			1.14			1.09			0			0			0
2132			Cu (%)			0.05			1.98			0.05			1.97			0			0			-0.01
2132			Fe (%)			11.08			1.22			11.08			1.22			0			0			0
2203			Ag (g/t)			126.03			1.48			105.84			1.21			-20.18			-0.16			-0.28
2203			Au (g/t)			0.65			1.63			0.65			1.63			0			0			0
2203			Pb (%)			1.46			0.76			1.46			0.76			0			0			0
2203			Zn (%)			5.1			1.11			5.1			1.11			0			0			0
2203			Cu (%)			0.02			0.62			0.02			0.7			0			0.13			0.07
2203			Fe (%)			20.8			0.82			20.8			0.82			0			0			0
2207			Ag (g/t)			131.52			1.62			131.52			1.62			0			0			0
2207			Au (g/t)			3.82			1.68			3.82			1.68			0			0			0
2207			Pb (%)			1.12			1.36			1.12			1.36			0			0			0
2207			Zn (%)			1.13			0.82			1.13			0.82			0			0			0
2207			Cu (%)			0.03			0.98			0.03			0.91			0			0.04			-0.07
2207			Fe (%)			19.12			0.67			19.12			0.67			0			0			0
2213			Ag (g/t)			83.75			0.96			83.75			0.96			0			0			0
2213			Au (g/t)			4.74			0.75			4.74			0.75			0			0			0
2213			Pb (%)			0.75			0.82			0.75			0.82			0			0			0
2213			Zn (%)			1.57			1.26			1.57			1.26			0			0			0
2213			Cu (%)			0.14			0.59			0.14			0.6			0			0.01			0.01
2213			Fe (%)			26.85			0.43			26.85			0.43			0			0			0
2215			Ag (g/t)			22.12			1.28			22.12			1.28			0			0			0
2215			Au (g/t)			0.43			1.07			0.43			1.07			0			0			0
2215			Pb (%)			1.22			1.2			1.22			1.2			0			0			0
2215			Zn (%)			16.14			1.49			16.14			1.49			0			0			0
2215			Cu (%)			0.05			1.18			0.05			1.18			0			0			0
2215			Fe (%)			39.98			0.88			39.98			0.88			0			0			0
2216			Ag (g/t)			28.53			0.82			28.53			0.82			0			0			0
2216			Au (g/t)			0.85			1.37			0.85			1.37			0			0			0
2216			Pb (%)			0.51			1.17			0.51			1.17			0			0			0
2216			Zn (%)			1.96			1.62			1.96			1.62			0			0			0
2216			Cu (%)			0.02			0.76			0.02			0.71			0			0.03			-0.04
2216			Fe (%)			13.51			0.47			13.51			0.47			0			0			0
2219			Ag (g/t)			159.53			1.36			159.53			1.36			0			0			0
2219			Au (g/t)			5.8			1.34			5.8			1.34			0			0			0
2219			Pb (%)			1.72			0.71			1.72			0.71			0			0			0
2219			Zn (%)			1.76			1.63			1.76			1.63			0			0			0
2219			Cu (%)			0.25			1.3			0.25			1.3			0			0			0
2219			Fe (%)			46.91			1.02			46.91			1.02			0			0			0
2221			Ag (g/t)			34.16			0.75			34.16			0.75			0			0			0
2221			Au (g/t)			0.69			1.37			0.69			1.37			0			0			0
2221			Pb (%)			1.05			0.88			1.05			0.88			0			0			0
2221			Zn (%)			1.07			1.01			1.07			1.01			0			0			0

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        455

---

2221			Cu (%)			0.03			1.17			0.03			1.15			0			0.01			-0.02
2221			Fe (%)			14.05			0.89			14.05			0.89			0			0			0
2222			Ag (g/t)			35.97			0.84			35.97			0.84			0			0			0
2222			Au (g/t)			0.31			0.48			0.31			0.48			0			0			0
2222			Pb (%)			1.84			1.18			1.84			1.18			0			0			0
2222			Zn (%)			20.38			1.39			16.46			1.28			-3.91			-0.19			-0.11
2222			Cu (%)			0.15			1.45			0.15			1.43			0			0.01			-0.02
2222			Fe (%)			12.75			0.45			12.75			0.45			0			0			0
2223			Ag (g/t)			105.52			1.35			95.32			1.25			-10.2			-0.1			-0.1
2223			Au (g/t)			2.77			1.56			2.77			1.56			0			0			0
2223			Pb (%)			1.15			0.92			1.15			0.92			0			0			0
2223			Zn (%)			3.84			1.17			3.84			1.17			0			0			0
2223			Cu (%)			0.04			0.87			0.04			0.85			0			0.05			-0.02
2223			Fe (%)			27.96			1.22			27.96			1.22			0			0			0
2224			Ag (g/t)			46.72			0.32			46.72			0.32			0			0			0
2224			Au (g/t)			1.41			1			1.41			1			0			0			0
2224			Pb (%)			5.15			0.91			2.6			0.81			-2.55			-0.5			-0.09
2224			Zn (%)			1.23			0.84			1.23			0.84			0			0			0
2224			Cu (%)			0.03			0.24			0.04			0.17			0			0.06			-0.07
2224			Fe (%)			12.9			0.78			12.9			0.78			0			0			0
2225			Ag (g/t)			44.93			1.47			44.93			1.47			0			0			0
2225			Au (g/t)			3.86			1.31			3.86			1.31			0			0			0
2225			Pb (%)			0.39			1.38			0.39			1.38			0			0			0
2225			Zn (%)			0.49			1.11			0.49			1.11			0			0			0
2225			Cu (%)			0.1			1.47			0.1			1.47			0			0			0
2225			Fe (%)			26.53			1.22			26.53			1.22			0			0			0
2235			Ag (g/t)			19.09			1.11			19.09			1.11			0			0			0
2235			Au (g/t)			4.07			1.52			4.07			1.52			0			0			0
2235			Pb (%)			0.56			1.22			0.56			1.22			0			0			0
2235			Zn (%)			3.28			0.78			3.28			0.78			0			0			0
2235			Cu (%)			0.03			0.58			0.03			0.58			0			0			0
2235			Fe (%)			14.27			1.03			14.27			1.03			0			0			0
2236			Ag (g/t)			106.32			0.68			106.32			0.68			0			0			0
2236			Au (g/t)			1.03			0.79			1.03			0.79			0			0			0
2236			Pb (%)			0.85			0.59			0.85			0.59			0			0			0
2236			Zn (%)			2.29			0.69			2.29			0.69			0			0			0
2236			Cu (%)			0.04			0.8			0.04			0.79			0			0.01			-0.02
2236			Fe (%)			27.33			0.37			27.33			0.37			0			0			0
2309			Ag (g/t)			117.84			1.65			117.84			1.65			0			0			0
2309			Au (g/t)			1.65			1.51			1.65			1.51			0			0			0
2309			Pb (%)			1.47			1.43			1.47			1.43			0			0			0
2309			Zn (%)			0.62			0.98			0.62			0.98			0			0			0
2309			Cu (%)			0.11			1.15			0.11			1.15			0			0			0
2309			Fe (%)			23.55			0.61			23.55			0.61			0			0			0
2311			Ag (g/t)			46.16			0.78			46.16			0.78			0			0			0
2311			Au (g/t)			1.72			0.85			1.72			0.85			0			0			0
2311			Pb (%)			1.07			1.72			1.07			1.72			0			0			0
2311			Zn (%)			3.22			1.39			3.22			1.39			0			0			0
2311			Cu (%)			0.07			1.22			0.07			1.21			0			0.01			-0.01
2311			Fe (%)			24.58			0.61			24.58			0.61			0			0			0
2314			Ag (g/t)			74.95			1.27			74.95			1.27			0			0			0

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        456

---

2314			Au (g/t)			0.68			0.57			0.68			0.57			0			0			0
2314			Pb (%)			1.06			0.78			1.06			0.78			0			0			0
2314			Zn (%)			2.82			1.34			2.82			1.34			0			0			0
2314			Cu (%)			0.16			0.74			0.16			0.74			0			0			0
2314			Fe (%)			28			0.65			28			0.65			0			0			0
2327			Ag (g/t)			33.13			0.55			33.13			0.55			0			0			0
2327			Au (g/t)			0.88			0.99			0.88			0.99			0			0			0
2327			Pb (%)			0.24			0.29			0.24			0.29			0			0			0
2327			Zn (%)			1.03			0.67			1.03			0.67			0			0			0
2327			Cu (%)			0.02			0.77			0.02			0.77			0			0.01			0
2327			Fe (%)			16.08			0.75			16.08			0.75			0			0			0
2400			Ag (g/t)			144.66			0.89			142.39			0.89			-2.27			-0.02			0
2400			Au (g/t)			2.76			1			2.76			1			0			0			0
2400			Pb (%)			1.34			2.11			1.11			1.68			-0.24			-0.17			-0.43
2400			Zn (%)			1.85			1.41			1.85			1.41			0			0			0
2400			Cu (%)			0.12			1.47			0.12			1.45			0			0.01			-0.02
2400			Fe (%)			41.14			0.7			41.14			0.7			0			0			0
2401			Ag (g/t)			43.93			1.02			43.93			1.02			0			0			0
2401			Au (g/t)			0.44			2.36			0.44			2.36			0			0			0
2401			Pb (%)			2.58			1.68			2.55			1.68			-0.03			-0.01			0
2401			Zn (%)			1.87			1.48			1.87			1.48			0			0			0
2401			Cu (%)			0.05			1.64			0.05			1.63			0			0			-0.01
2401			Fe (%)			15.4			1.68			15.4			1.68			0			0			0
2405			Ag (g/t)			134.01			1.43			51.73			0.64			-82.28			-0.61			-0.79
2405			Au (g/t)			1.81			1.58			0.56			0.79			-1.25			-0.69			-0.79
2405			Pb (%)			1.6			0.86			1.6			0.86			0			0			0
2405			Zn (%)			3.33			1.16			2.27			0.94			-1.06			-0.32			-0.22
2405			Cu (%)			0.05			0.58			0.06			0.6			0			0.04			0.01
2405			Fe (%)			55.31			1.12			48.07			0.99			-7.24			-0.13			-0.12
2408			Ag (g/t)			69.81			1.17			69.81			1.17			0			0			0
2408			Au (g/t)			1.03			0.66			1.03			0.66			0			0			0
2408			Pb (%)			0.47			1.41			0.47			1.41			0			0			0
2408			Zn (%)			0.69			1.38			0.69			1.38			0			0			0
2408			Cu (%)			0.09			1.75			0.09			1.75			0			0			0
2408			Fe (%)			13.43			0.54			13.43			0.54			0			0			0
2410			Ag (g/t)			117.02			1.79			119.56			1.8			2.54			0.02			0.01
2410			Au (g/t)			2.39			2.03			2.43			2.03			0.04			0.02			0
2410			Pb (%)			1.4			1.23			1.4			1.23			0			0			0
2410			Zn (%)			3.33			1.07			3.36			1.07			0.03			0.01			0
2410			Cu (%)			0.08			1.98			0.08			1.96			0			0.01			-0.02
2410			Fe (%)			30.62			1.23			30.62			1.23			0			0			0
2411			Ag (g/t)			29.57			0.81			29.57			0.81			0			0			0
2411			Au (g/t)			2.3			1.34			2.3			1.34			0			0			0
2411			Pb (%)			0.34			0.84			0.34			0.84			0			0			0
2411			Zn (%)			0.75			0.81			0.75			0.81			0			0			0
2411			Cu (%)			0.05			0.86			0.05			0.86			0			0			0
2411			Fe (%)			34.54			1.13			34.54			1.13			0			0			0
2417			Ag (g/t)			33.86			0.84			33.86			0.84			0			0			0
2417			Au (g/t)			1.34			1.01			1.34			1.01			0			0			0
2417			Pb (%)			0.33			0.99			0.33			0.99			0			0			0
2417			Zn (%)			0.59			1.07			0.59			1.07			0			0			0

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        457

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2417			Cu (%)			0.01			0.35			0.01			0.35			0			0			0
2417			Fe (%)			20.67			0.61			20.67			0.61			0			0			0
2424			Ag (g/t)			21.66			0.42			21.66			0.42			0			0			0
2424			Au (g/t)			0.59			0.84			0.59			0.84			0			0			0
2424			Pb (%)			0.59			0.8			0.59			0.8			0			0			0
2424			Zn (%)			1.44			0.78			1.44			0.78			0			0			0
2424			Cu (%)			0.02			0.89			0.02			0.89			0			0			0
2424			Fe (%)			15.1			0.51			15.1			0.51			0			0			0
2425			Ag (g/t)			26.48			0.77			26.48			0.77			0			0			0
2425			Au (g/t)			0.2			1.66			0.2			1.66			0			0			0
2425			Pb (%)			1.73			1.26			1.65			1.24			-0.08			-0.05			-0.02
2425			Zn (%)			0.92			1.11			0.92			1.11			0			0			0
2425			Cu (%)			0.21			1.08			0.21			1.08			0			0			0
2425			Fe (%)			2.2			0.74			2.2			0.74			0			0			0
2502			Ag (g/t)			42.23			1.12			42.23			1.12			0			0			0
2502			Au (g/t)			2.15			1.39			2.15			1.39			0			0			0
2502			Pb (%)			1.61			1.64			1.47			1.6			-0.15			-0.09			-0.04
2502			Zn (%)			1.45			1.36			1.45			1.36			0			0			0
2502			Cu (%)			0.2			1.44			0.2			1.44			0			0			0
2502			Fe (%)			34.42			1.01			34.42			1.01			0			0			0
2507			Ag (g/t)			79.07			1.57			79.07			1.57			0			0			0
2507			Au (g/t)			2.82			2.87			2.82			2.87			0			0			0
2507			Pb (%)			0.88			1.5			0.88			1.5			0			0			0
2507			Zn (%)			2.19			1.56			2.19			1.56			0			0			0
2507			Cu (%)			0.12			1.67			0.12			1.65			0			0.01			-0.02
2507			Fe (%)			21.73			1.39			21.73			1.39			0			0			0
2510			Ag (g/t)			48.26			0.83			48.26			0.83			0			0			0
2510			Au (g/t)			1.28			1.24			1.28			1.24			0			0			0
2510			Pb (%)			0.71			0.83			0.71			0.83			0			0			0
2510			Zn (%)			0.89			0.84			0.89			0.84			0			0			0
2510			Cu (%)			0.24			1.26			0.24			1.26			0			0			0
2510			Fe (%)			53.06			1.33			46.81			1.2			-6.25			-0.12			-0.13
2538			Ag (g/t)			12.49			0.73			12.49			0.73			0			0			0
2538			Au (g/t)			0.1			0.97			0.1			0.97			0			0			0
2538			Pb (%)			0.8			1.47			0.8			1.47			0			0			0
2538			Zn (%)			3.1			1.97			2.85			1.9			-0.25			-0.08			-0.07
2538			Cu (%)			0.02			1.5			0.02			1.44			0			0.03			-0.06
2538			Fe (%)			6.19			0.55			6.19			0.55			0			0			0
2614			Ag (g/t)			51.77			0.59			51.77			0.59			0			0			0
2614			Au (g/t)			0.4			0.83			0.4			0.83			0			0			0
2614			Pb (%)			0.96			0.72			0.96			0.72			0			0			0
2614			Zn (%)			3.33			1.04			3.33			1.04			0			0			0
2614			Cu (%)			0.27			1.67			0.27			1.66			0			0.01			-0.01
2614			Fe (%)			35.15			0.53			35.15			0.53			0			0			0
2729			Ag (g/t)			36.32			0.72			36.32			0.72			0			0			0
2729			Au (g/t)			1.87			0.82			1.87			0.82			0			0			0
2729			Pb (%)			0.51			0.28			0.51			0.28			0			0			0
2729			Zn (%)			0.61			0.51			0.61			0.51			0			0			0
2729			Cu (%)			0.09			1.06			0.09			1.06			0			0			0
2729			Fe (%)			32.21			0.72			32.21			0.72			0			0			0
2802			Ag (g/t)			63.43			1.1			62.05			1.08			-1.38			-0.02			-0.02

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        458

---

2802			Au (g/t)			1.36			1.32			1.36			1.32			0			0			0
2802			Pb (%)			2.56			1.12			2.56			1.12			0			0			0
2802			Zn (%)			4.83			1.39			4.83			1.39			0			0			0
2802			Cu (%)			0.05			1.08			0.05			1.06			0			0.01			-0.02
2802			Fe (%)			16.78			1.07			16.78			1.07			0			0			0
3001			Ag (g/t)			68.96			2.07			49.24			1.29			-19.72			-0.29			-0.78
3001			Au (g/t)			1.34			1.49			1.34			1.49			0			0			0
3001			Pb (%)			0.64			0.97			0.64			0.97			0			0			0
3001			Zn (%)			1.62			1.21			1.62			1.21			0			0			0
3001			Cu (%)			0.1			1.67			0.1			1.67			0			0			0
3001			Fe (%)			28.67			1.16			28.67			1.16			0			0			0
3002			Ag (g/t)			28.9			0.54			28.9			0.54			0			0			0
3002			Au (g/t)			0.72			0.52			0.72			0.52			0			0			0
3002			Pb (%)			0.76			1.33			0.76			1.33			0			0			0
3002			Zn (%)			1.5			0.82			1.5			0.82			0			0			0
3002			Cu (%)			0.02			0.5			0.03			0.5			0			0.01			0
3002			Fe (%)			14.06			0.45			14.06			0.45			0			0			0
3003			Ag (g/t)			17.77			0.93			17.77			0.93			0			0			0
3003			Au (g/t)			0.19			0.85			0.19			0.85			0			0			0
3003			Pb (%)			0.39			0.84			0.39			0.84			0			0			0
3003			Zn (%)			0.78			1.05			0.78			1.05			0			0			0
3003			Cu (%)			0.04			0.56			0.04			0.56			0			0			0
3003			Fe (%)			16.71			0.67			16.71			0.67			0			0			0
3004			Ag (g/t)			31.97			0.93			31.97			0.93			0			0			0
3004			Au (g/t)			1.19			1.29			1.19			1.29			0			0			0
3004			Pb (%)			0.85			1.79			0.85			1.79			0			0			0
3004			Zn (%)			1.34			1.09			1.34			1.09			0			0			0
3004			Cu (%)			0.05			1.37			0.05			1.38			0			0.02			0.02
3004			Fe (%)			16.74			0.77			16.74			0.77			0			0			0
3005			Ag (g/t)			61.28			0.68			61.28			0.68			0			0			0
3005			Au (g/t)			2.51			1.2			2.51			1.2			0			0			0
3005			Pb (%)			0.83			0.7			0.83			0.7			0			0			0
3005			Zn (%)			1.91			1.29			1.91			1.29			0			0			0
3005			Cu (%)			0.05			0.75			0.05			0.74			0			0.02			-0.01
3005			Fe (%)			35.41			0.76			35.41			0.76			0			0			0
3006			Ag (g/t)			55.43			1.18			55.43			1.18			0			0			0
3006			Au (g/t)			1.18			0.98			1.18			0.98			0			0			0
3006			Pb (%)			0.85			0.92			0.85			0.92			0			0			0
3006			Zn (%)			5.2			0.97			4.66			0.9			-0.53			-0.1			-0.07
3006			Cu (%)			0.15			1.37			0.15			1.37			0			0			0
3006			Fe (%)			35.87			1.03			35.87			1.03			0			0			0
3007			Ag (g/t)			56.4			1.07			56.4			1.07			0			0			0
3007			Au (g/t)			0.49			1.48			0.49			1.48			0			0			0
3007			Pb (%)			0.9			1.21			0.9			1.21			0			0			0
3007			Zn (%)			2			0.44			2			0.44			0			0			0
3007			Cu (%)			0.12			1.06			0.12			1.06			0			0			0
3007			Fe (%)			27.03			0.55			27.03			0.55			0			0			0
3008			Ag (g/t)			83.32			0.85			83.32			0.85			0			0			0
3008			Au (g/t)			0.32			0.53			0.32			0.53			0			0			0
3008			Pb (%)			0.56			0.61			0.56			0.61			0			0			0
3008			Zn (%)			1.94			0.45			1.94			0.45			0			0			0

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        459

---

3008			Cu (%)			0.04			0.55			0.04			0.55			0			0			0
3008			Fe (%)			40.74			0.97			40.74			0.97			0			0			0
3009			Ag (g/t)			90.82			1.49			90.82			1.49			0			0			0
3009			Au (g/t)			6.23			2.56			6.23			2.56			0			0			0
3009			Pb (%)			1.2			1.17			1.2			1.17			0			0			0
3009			Zn (%)			1.71			0.9			1.71			0.9			0			0			0
3009			Cu (%)			0.12			1.66			0.12			1.65			0			0.01			0
3009			Fe (%)			57.44			0.99			57.44			0.99			0			0			0
3010			Ag (g/t)			29.98			1.27			29.98			1.27			0			0			0
3010			Au (g/t)			2.12			1.01			2.12			1.01			0			0			0
3010			Pb (%)			0.5			1.09			0.5			1.09			0			0			0
3010			Zn (%)			0.61			0.94			0.61			0.94			0			0			0
3010			Cu (%)			0.1			1.63			0.1			1.63			0			0			0
3010			Fe (%)			18.53			0.77			18.53			0.77			0			0			0
3011			Ag (g/t)			20.77			0.66			20.77			0.66			0			0			0
3011			Au (g/t)			0.64			0.59			0.64			0.59			0			0			0
3011			Pb (%)			0.66			0.95			0.66			0.95			0			0			0
3011			Zn (%)			2.56			0.74			2.55			0.74			-0.01			-0.01			0
3011			Cu (%)			0.03			0.77			0.03			0.76			0			0.01			-0.01
3011			Fe (%)			17.48			0.77			16.37			0.67			-1.1			-0.06			-0.1
3100			Ag (g/t)			13.19			0.73			13.19			0.73			0			0			0
3100			Au (g/t)			0.11			0.34			0.11			0.34			0			0			0
3100			Pb (%)			0.48			0.75			0.48			0.75			0			0			0
3100			Zn (%)			3.46			0.89			3.34			0.89			-0.13			-0.04			0
3100			Cu (%)			0.01			0.52			0.01			0.59			0			0.17			0.07
3100			Fe (%)			4.96			0.67			4.96			0.67			0			0			0
3101			Ag (g/t)			58.14			1.34			54.41			1.14			-3.73			-0.06			-0.2
3101			Au (g/t)			2.26			1.31			2.18			1.2			-0.08			-0.04			-0.1
3101			Pb (%)			0.64			1.66			0.61			1.57			-0.02			-0.03			-0.08
3101			Zn (%)			1.58			1.32			1.49			1.19			-0.09			-0.06			-0.13
3101			Cu (%)			0.14			1.99			0.12			1.63			-0.01			-0.11			-0.36
3101			Fe (%)			30.86			1.08			30.86			1.08			0			0			0
3102			Ag (g/t)			107.27			1.43			99.83			1.32			-7.43			-0.07			-0.11
3102			Au (g/t)			1.94			1.07			1.94			1.07			0			0			0
3102			Pb (%)			1.06			1.57			1.06			1.57			0			0			0
3102			Zn (%)			1.91			1.24			1.9			1.23			-0.01			0			-0.01
3102			Cu (%)			0.11			1.24			0.11			1.24			0			0			0
3102			Fe (%)			31.26			0.75			31.26			0.75			0			0			0
3103			Ag (g/t)			59.16			1.56			52.54			1.34			-6.62			-0.11			-0.21
3103			Au (g/t)			3.32			1.53			2.56			1.33			-0.77			-0.23			-0.2
3103			Pb (%)			0.35			0.86			0.35			0.86			0			0			0
3103			Zn (%)			1.28			1.89			0.84			0.88			-0.44			-0.34			-1.02
3103			Cu (%)			0.18			2.35			0.16			2.2			-0.02			-0.14			-0.16
3103			Fe (%)			34.74			1.2			34.74			1.2			0			0			0
3121			Ag (g/t)			32.97			0.8			32.97			0.8			0			0			0
3121			Au (g/t)			1.64			1.55			1.64			1.55			0			0			0
3121			Pb (%)			0.36			0.96			0.36			0.96			0			0			0
3121			Zn (%)			0.87			0.95			0.87			0.95			0			0			0
3121			Cu (%)			0.11			1.33			0.11			1.33			0			0			-0.01
3121			Fe (%)			25.84			1.17			25.05			1.11			-0.79			-0.03			-0.05
3191			Ag (g/t)			74.14			1.43			74.14			1.43			0			0			0

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        460

---

3191			Au (g/t)			3.79			1.49			3.79			1.49			0			0			0
3191			Pb (%)			0.54			1.38			0.54			1.38			0			0			0
3191			Zn (%)			1.97			1.65			1.97			1.65			0			0			0
3191			Cu (%)			0.15			1.93			0.15			1.93			0			0			-0.01
3191			Fe (%)			35.56			1.1			35.56			1.1			0			0			0
3192			Ag (g/t)			65.74			1.15			65.74			1.15			0			0			0
3192			Au (g/t)			2.24			0.95			2.24			0.95			0			0			0
3192			Pb (%)			0.59			1.55			0.59			1.55			0			0			0
3192			Zn (%)			1.57			0.96			1.57			0.96			0			0			0
3192			Cu (%)			0.11			1.34			0.11			1.34			0			0			0
3192			Fe (%)			31.29			0.63			31.29			0.63			0			0			0
3193			Ag (g/t)			35.42			1.39			35.42			1.39			0			0			0
3193			Au (g/t)			2.38			2.27			2.38			2.27			0			0			0
3193			Pb (%)			0.34			0.81			0.34			0.81			0			0			0
3193			Zn (%)			1			1.02			1			1.02			0			0			0
3193			Cu (%)			0.05			1.19			0.05			1.18			0			0.01			-0.01
3193			Fe (%)			24.02			1.08			24.02			1.08			0			0			0
3204			Ag (g/t)			41.23			0.82			41.23			0.82			0			0			0
3204			Au (g/t)			1.18			1.06			1.18			1.06			0			0			0
3204			Pb (%)			0.36			1.36			0.36			1.36			0			0			0
3204			Zn (%)			1.01			1.56			1.01			1.56			0			0			0
3204			Cu (%)			0.07			1.79			0.07			1.78			0			0.01			-0.01
3204			Fe (%)			18.04			0.82			18.04			0.82			0			0			0
3205			Ag (g/t)			34.21			2.66			34.21			2.66			0			0			0
3205			Au (g/t)			0.23			1.2			0.23			1.2			0			0			0
3205			Pb (%)			0.31			1.04			0.31			1.04			0			0			0
3205			Zn (%)			0.98			0.83			0.98			0.83			0			0			0
3205			Cu (%)			0.03			0.83			0.03			0.83			0			0.01			0
3205			Fe (%)			13.45			0.82			13.45			0.82			0			0			0
3206			Ag (g/t)			32.28			1.01			32.28			1.01			0			0			0
3206			Au (g/t)			1.03			1.12			1.03			1.12			0			0			0
3206			Pb (%)			0.49			1.34			0.49			1.34			0			0			0
3206			Zn (%)			1.73			1.5			1.73			1.5			0			0			0
3206			Cu (%)			0.07			0.99			0.07			1			0			0.01			0.01
3206			Fe (%)			23.73			0.74			23.73			0.74			0			0			0
3207			Ag (g/t)			10.94			0.44			10.94			0.44			0			0			0
3207			Au (g/t)			0.59			1			0.59			1			0			0			0
3207			Pb (%)			0.22			1.31			0.22			1.31			0			0			0
3207			Zn (%)			0.45			1.05			0.45			1.05			0			0			0
3207			Cu (%)			0.05			0.88			0.05			0.88			0			0			0
3207			Fe (%)			15.13			0.53			15.13			0.53			0			0			0
3222			Ag (g/t)			20.1			1.2			20.1			1.2			0			0			0
3222			Au (g/t)			0.62			2.99			0.54			2.8			-0.08			-0.13			-0.19
3222			Pb (%)			0.57			1.53			0.57			1.53			0			0			0
3222			Zn (%)			1.16			1.12			1.16			1.12			0			0			0
3222			Cu (%)			0.07			1.8			0.07			1.8			0			0			0
3222			Fe (%)			11.72			1.13			11.72			1.13			0			0			0
3223			Ag (g/t)			31.2			1.36			31.2			1.36			0			0			0
3223			Au (g/t)			1.03			1.74			1.03			1.74			0			0			0
3223			Pb (%)			1.33			1.56			1.33			1.56			0			0			0
3223			Zn (%)			2.4			1.57			2.38			1.56			-0.01			-0.01			0

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        461

---

3223			Cu (%)			0.07			2.04			0.07			2.02			0			0.01			-0.02
3223			Fe (%)			18			1.1			18			1.1			0			0			0
3224			Ag (g/t)			20.79			1.25			17.74			0.92			-3.05			-0.15			-0.33
3224			Au (g/t)			0.37			1.68			0.37			1.68			0			0			0
3224			Pb (%)			0.43			1.02			0.43			1.02			0			0			0
3224			Zn (%)			1.08			0.77			1.08			0.77			0			0			0
3224			Cu (%)			0.1			2.39			0.1			2.39			0			0			0
3224			Fe (%)			11.04			0.38			11.04			0.38			0			0			0
3225			Ag (g/t)			51.64			0.61			51.64			0.61			0			0			0
3225			Au (g/t)			2.2			0.84			2.2			0.84			0			0			0
3225			Pb (%)			3.52			0.85			3.52			0.85			0			0			0
3225			Zn (%)			2.44			0.54			2.44			0.54			0			0			0
3225			Cu (%)			0.29			0.6			0.29			0.6			0			0			0
3225			Fe (%)			30.78			0.56			30.78			0.56			0			0			0
3226			Ag (g/t)			29.25			1.6			29.25			1.6			0			0			0
3226			Au (g/t)			0.69			1.9			0.62			1.75			-0.07			-0.1			-0.15
3226			Pb (%)			0.96			1.09			0.96			1.09			0			0			0
3226			Zn (%)			1.87			1.56			1.82			1.58			-0.05			-0.03			0.01
3226			Cu (%)			0.11			1.5			0.11			1.5			0			0			0
3226			Fe (%)			16.46			1.48			16.46			1.48			0			0			0
3227			Ag (g/t)			59.16			0.87			59.16			0.87			0			0			0
3227			Au (g/t)			1.74			0.96			1.74			0.96			0			0			0
3227			Pb (%)			0.62			1.35			0.62			1.35			0			0			0
3227			Zn (%)			1.72			1.05			1.72			1.05			0			0			0
3227			Cu (%)			0.09			0.98			0.09			0.97			0			0.01			-0.01
3227			Fe (%)			21.58			0.54			21.58			0.54			0			0			0
3230			Ag (g/t)			114.03			1.07			63.05			0.8			-50.98			-0.45			-0.28
3230			Au (g/t)			0.79			0.73			0.79			0.73			0			0			0
3230			Pb (%)			1.33			0.24			1.33			0.24			0			0			0
3230			Zn (%)			1.84			0.39			1.84			0.39			0			0			0
3230			Cu (%)			0.05			0.15			0.05			0.13			0			0.03			-0.03
3230			Fe (%)			17.46			0.18			17.46			0.18			0			0			0
3291			Ag (g/t)			104.12			1.22			103.52			1.21			-0.6			-0.01			-0.01
3291			Au (g/t)			2.49			1.63			2.28			1.23			-0.21			-0.09			-0.39
3291			Pb (%)			1.55			1.61			1.47			1.46			-0.08			-0.05			-0.15
3291			Zn (%)			2.9			1.55			2.68			1.29			-0.22			-0.07			-0.26
3291			Cu (%)			0.13			1.48			0.13			1.24			-0.01			-0.05			-0.25
3291			Fe (%)			43.74			0.86			42.7			0.81			-1.04			-0.02			-0.04
3292			Ag (g/t)			56.92			0.95			56.92			0.95			0			0			0
3292			Au (g/t)			1.86			1.1			1.75			0.98			-0.11			-0.06			-0.13
3292			Pb (%)			0.81			1.32			0.72			1.1			-0.09			-0.11			-0.22
3292			Zn (%)			2.02			1.04			2.02			1.04			0			0			0
3292			Cu (%)			0.09			1.01			0.08			0.85			-0.01			-0.06			-0.15
3292			Fe (%)			36.06			0.78			36.06			0.78			0			0			0
3293			Ag (g/t)			31.54			1.05			31.54			1.05			0			0			0
3293			Au (g/t)			0.71			1.11			0.71			1.11			0			0			0
3293			Pb (%)			0.75			1.21			0.75			1.21			0			0			0
3293			Zn (%)			1.46			0.95			1.46			0.95			0			0			0
3293			Cu (%)			0.05			1.25			0.05			1.25			0			0.01			-0.01
3293			Fe (%)			21.76			0.63			21.76			0.63			0			0			0
3294			Ag (g/t)			90.46			0.94			90.46			0.94			0			0			0

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3294			Au (g/t)			3.81			1.21			3.81			1.21			0			0			0
3294			Pb (%)			0.11			0.65			0.11			0.65			0			0			0
3294			Zn (%)			0.49			1.24			0.49			1.24			0			0			0
3294			Cu (%)			0.13			0.5			0.13			0.5			0			0			0
3294			Fe (%)			80.77			0.88			48.25			0.62			-32.51			-0.4			-0.26
3295			Ag (g/t)			38.4			1.29			38.4			1.29			0			0			0
3295			Au (g/t)			0.38			1.09			0.38			1.09			0			0			0
3295			Pb (%)			0.78			1.35			0.78			1.35			0			0			0
3295			Zn (%)			1.47			1.19			1.47			1.19			0			0			0
3295			Cu (%)			0.1			1.77			0.1			1.77			0			0			0
3295			Fe (%)			23.8			1.69			22.55			1.49			-1.25			-0.05			-0.2
3301			Ag (g/t)			234.86			4.14			130.84			2.14			- 104.02			-0.44			-2
3301			Au (g/t)			1.3			1.49			1.2			1.36			-0.1			-0.08			-0.12
3301			Pb (%)			0.95			1.35			0.92			1.28			-0.03			-0.03			-0.07
3301			Zn (%)			1.7			1.26			1.7			1.26			0			0			0
3301			Cu (%)			0.2			1.96			0.19			1.78			-0.01			-0.06			-0.18
3301			Fe (%)			28.04			0.96			26.62			0.84			-1.42			-0.05			-0.12
3402			Ag (g/t)			13.83			0.35			13.83			0.35			0			0			0
3402			Au (g/t)			0.41			0.89			0.41			0.89			0			0			0
3402			Pb (%)			0.24			0.98			0.24			0.98			0			0			0
3402			Zn (%)			0.7			0.8			0.7			0.8			0			0			0
3402			Cu (%)			0.03			0.44			0.03			0.43			0			0.01			-0.01
3402			Fe (%)			16.67			0.63			16.67			0.63			0			0			0
3409			Ag (g/t)			22.22			0.91			22.22			0.91			0			0			0
3409			Au (g/t)			1.03			0.77			1.03			0.77			0			0			0
3409			Pb (%)			0.42			1.25			0.42			1.25			0			0			0
3409			Zn (%)			1.4			0.99			1.4			0.99			0			0			0
3409			Cu (%)			0.02			1.17			0.02			1.19			0			0.01			0.02
3409			Fe (%)			11.68			0.76			11.68			0.76			0			0			0
3410			Ag (g/t)			11.81			0.98			11.81			0.98			0			0			0
3410			Au (g/t)			0.64			1.49			0.64			1.49			0			0			0
3410			Pb (%)			0.1			0.71			0.1			0.71			0			0			0
3410			Zn (%)			0.73			0.96			0.73			0.96			0			0			0
3410			Cu (%)			0.02			0.56			0.02			0.56			0			0.01			0.01
3410			Fe (%)			14.16			0.93			14.16			0.93			0			0			0
3411			Ag (g/t)			28.72			0.73			28.72			0.73			0			0			0
3411			Au (g/t)			1.24			0.87			1.2			0.82			-0.04			-0.03			-0.05
3411			Pb (%)			1			0.82			1			0.82			0			0			0
3411			Zn (%)			2.46			0.9			2.46			0.9			0			0			0
3411			Cu (%)			0.05			0.97			0.05			0.96			0			0.01			-0.01
3411			Fe (%)			22.58			0.73			22.58			0.73			0			0			0
3508			Ag (g/t)			22.47			0.77			22.47			0.77			0			0			0
3508			Au (g/t)			0.87			0.69			0.87			0.69			0			0			0
3508			Pb (%)			0.63			0.97			0.63			0.97			0			0			0
3508			Zn (%)			1.68			0.84			1.63			0.8			-0.05			-0.03			-0.04
3508			Cu (%)			0.02			0.47			0.02			0.48			0			0.02			0.01
3508			Fe (%)			16.79			0.5			16.79			0.5			0			0			0
3600			Ag (g/t)			30.37			0.72			29.73			0.72			-0.64			-0.02			0
3600			Au (g/t)			0.72			0.97			0.72			0.98			0			0			0
3600			Pb (%)			0.88			0.95			0.88			0.95			0			0			0
3600			Zn (%)			2.59			0.96			2.59			0.96			0			0			0

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        463

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3600			Cu (%)			0.04			1.48			0.04			1.47			0			0.01			-0.02
3600			Fe (%)			16.08			0.84			16.08			0.84			0			0			0
4001			Ag (g/t)			924.75			1.19			722.7			1.12			- 202.05			-0.22			-0.06
4001			Au (g/t)			0.07			0.59			0.07			0.59			0			0			0
4001			Pb (%)			0.21			0.78			0.21			0.78			0			0			0
4001			Zn (%)			0.16			0.57			0.16			0.57			0			0			0
4001			Cu (%)			2.92			0.22			2.92			0.22			0			0			0
4001			Fe (%)			6.94			0.44			6.94			0.44			0			0			0
4002			Ag (g/t)			39.79			1			39.79			1			0			0			0
4002			Au (g/t)			0.01			0.48			0.01			0.48			0			0			0
4002			Pb (%)			2.38			1.04			2.38			1.04			0			0			0
4002			Zn (%)			3.96			0.7			3.96			0.7			0			0			0
4002			Cu (%)			0.09			0.41			0.09			0.41			0			0			0
4002			Fe (%)			2.35			0.37			2.35			0.37			0			0			0
4003			Ag (g/t)			94.14			0.34			94.14			0.34			0			0			0
4003			Au (g/t)			0.07			0.06			0.07			0.06			0			0			0
4003			Pb (%)			4.52			0.26			4.52			0.26			0			0			0
4003			Zn (%)			23.75			0.83			23.75			0.83			0			0			0
4003			Cu (%)			0.3			0.03			0.3			0.03			0			0			0
4003			Fe (%)			3.84			0.7			3.84			0.7			0			0			0

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        464

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Appendix CSTATISTICAL VALIDATION _ CAPPED COMPOSITES VS BLOCK MODEL

Capped composites vs block model (back calculated assay grades)

Domain			Assay			Composite Count			Composite Mean			Block Model Mean			Diff (BM - Composite)			% Diff vs Composite
1003			Ag			9			25.006			43.452			18.446			0.738
1003			Au			9			0.842			0.65			-0.192			-0.228
1003			Cu			9			0.124			0.07			-0.054			-0.435
1003			Fe			9			15.167			21.307			6.141			0.405
1003			Pb			9			0.43			0.39			-0.041			-0.095
1003			Zn			9			1.087			1.65			0.563			0.518
1004			Ag			9			39.464			56.953			17.488			0.443
1004			Au			9			0.29			0.322			0.033			0.113
1004			Cu			9			0.043			0.035			-0.008			-0.179
1004			Fe			9			9.642			8.809			-0.833			-0.086
1004			Pb			9			1.156			1.174			0.018			0.016
1004			Zn			9			1.919			1.607			-0.312			-0.163
1005			Ag			4			72.631			82.245			9.614			0.132
1005			Au			4			0.579			0.584			0.005			0.009
1005			Cu			4			0.036			0.042			0.006			0.153
1005			Fe			4			11.318			11.652			0.335			0.03
1005			Pb			4			0.342			0.298			-0.044			-0.13
1005			Zn			4			1.494			1.083			-0.411			-0.275
1006			Ag			4			47.424			24.022			-23.402			-0.493
1006			Au			4			0.531			0.298			-0.233			-0.439
1006			Cu			4			0.026			0.018			-0.008			-0.304
1006			Fe			4			11.135			9.846			-1.289			-0.116
1006			Pb			4			0.7			0.356			-0.344			-0.492
1006			Zn			4			3.05			1.094			-1.956			-0.641
1007			Ag			5			190.479			117.16			-73.319			-0.385
1007			Au			5			0.489			0.434			-0.055			-0.113
1007			Cu			5			0.249			0.144			-0.106			-0.424
1007			Fe			5			10.196			9.15			-1.046			-0.103
1007			Pb			5			0.34			0.439			0.099			0.293
1007			Zn			5			2.006			1.845			-0.16			-0.08
1011			Ag			3			36.567			33.313			-3.254			-0.089
1011			Au			3			1.364			0.931			-0.433			-0.318
1011			Cu			3			0.032			0.021			-0.011			-0.352
1011			Fe			3			21.01			25.409			4.399			0.209
1011			Pb			3			0.329			0.274			-0.055			-0.166
1011			Zn			3			0.639			0.537			-0.102			-0.16
1012			Ag			2			63.015			68.283			5.268			0.084
1012			Au			2			0.533			0.607			0.074			0.139
1012			Cu			2			0.123			0.113			-0.01			-0.082
1012			Fe			2			11.265			12.505			1.241			0.11
1012			Pb			2			0.469			0.409			-0.06			-0.128
1012			Zn			2			0.837			0.708			-0.129			-0.154
1013			Ag			3			28.711			33.134			4.423			0.154
1013			Au			3			0.144			0.134			-0.01			-0.071

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1013			Cu			3			0.031			0.026			-0.005			-0.171
1013			Fe			3			15.41			20.177			4.767			0.309
1013			Pb			3			0.281			0.58			0.299			1.064
1013			Zn			3			0.586			1.277			0.692			1.18
1030			Ag			9			67.483			85.794			18.312			0.271
1030			Au			9			0.906			1.634			0.728			0.804
1030			Cu			9			0.093			0.158			0.065			0.696
1030			Fe			9			15.351			15.146			-0.205			-0.013
1030			Pb			9			1.051			1.08			0.029			0.027
1030			Zn			9			3.831			5.279			1.449			0.378
1031			Ag			31			79.26			61.836			-17.424			-0.22
1031			Au			31			0.586			1.21			0.624			1.064
1031			Cu			31			0.049			0.062			0.013			0.254
1031			Fe			31			13.762			19.455			5.693			0.414
1031			Pb			31			0.737			0.556			-0.182			-0.247
1031			Zn			31			1.666			2.035			0.369			0.222
1033			Ag			6			20.884			19.051			-1.833			-0.088
1033			Au			6			0.767			1.046			0.28			0.365
1033			Cu			6			0.035			0.023			-0.012			-0.342
1033			Fe			6			11.637			13.512			1.875			0.161
1033			Pb			6			0.34			0.218			-0.122			-0.359
1033			Zn			6			0.462			0.256			-0.206			-0.447
1040			Ag			11			48.208			36.631			-11.577			-0.24
1040			Au			11			1.05			1.282			0.232			0.22
1040			Cu			11			0.039			0.027			-0.011			-0.296
1040			Fe			11			12.626			12.015			-0.611			-0.048
1040			Pb			11			0.789			1.219			0.43			0.545
1040			Zn			11			1.24			2.325			1.084			0.874
1041			Ag			26			85.314			88.671			3.357			0.039
1041			Au			26			0.904			1.284			0.38			0.42
1041			Cu			26			0.03			0.032			0.002			0.053
1041			Fe			26			14.664			13.173			-1.491			-0.102
1041			Pb			26			1.332			1.354			0.022			0.017
1041			Zn			26			2.153			2.05			-0.102			-0.048
1050			Ag			11			23.867			35.727			11.859			0.497
1050			Au			11			1.132			2.034			0.903			0.798
1050			Cu			11			0.368			0.123			-0.245			-0.667
1050			Fe			11			11.143			18.397			7.253			0.651
1050			Pb			11			0.363			0.49			0.127			0.351
1050			Zn			11			1.175			1.284			0.109			0.093
1051			Ag			2			67.605			77.187			9.582			0.142
1051			Au			2			0.845			0.736			-0.109			-0.129
1051			Cu			2			0.03			0.034			0.004			0.122
1051			Fe			2			9.998			8.919			-1.079			-0.108
1051			Pb			2			0.271			0.301			0.03			0.109
1051			Zn			2			0.085			0.103			0.018			0.213
1052			Ag			31			46.61			50.742			4.132			0.089
1052			Au			31			0.992			0.502			-0.489			-0.493
1052			Cu			31			0.14			0.115			-0.025			-0.179
1052			Fe			31			13.861			14.403			0.542			0.039
1052			Pb			31			0.809			0.279			-0.53			-0.655

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1052			Zn			31			0.877			0.579			-0.298			-0.34
1053			Ag			6			51.842			28.546			-23.296			-0.449
1053			Au			6			0.823			0.65			-0.173			-0.211
1053			Cu			6			0.068			0.053			-0.015			-0.216
1053			Fe			6			12.348			10.279			-2.069			-0.168
1053			Pb			6			1.039			0.462			-0.577			-0.555
1053			Zn			6			3.246			1.251			-1.995			-0.615
1054			Ag			7			21.778			18.616			-3.162			-0.145
1054			Au			7			0.04			0.043			0.004			0.092
1054			Cu			7			0.459			0.194			-0.265			-0.577
1054			Fe			7			5.623			8.377			2.754			0.49
1054			Pb			7			0.654			0.808			0.154			0.235
1054			Zn			7			0.846			2.037			1.191			1.408
1055			Ag			2			202.533			189.429			-13.104			-0.065
1055			Au			2			0.479			0.463			-0.016			-0.034
1055			Cu			2			0.033			0.047			0.013			0.401
1055			Fe			2			13.436			12.766			-0.67			-0.05
1055			Pb			2			3.121			3.035			-0.086			-0.027
1055			Zn			2			2.363			2.911			0.549			0.232
1056			Ag			8			25.144			26.189			1.045			0.042
1056			Au			8			0.339			0.394			0.054			0.16
1056			Cu			8			0.468			0.247			-0.221			-0.472
1056			Fe			8			17.736			18.237			0.5			0.028
1056			Pb			8			0.369			0.364			-0.004			-0.012
1056			Zn			8			0.352			0.314			-0.038			-0.107
1061			Ag			2			50.719			56.54			5.82			0.115
1061			Au			2			0.236			0.292			0.055			0.234
1061			Cu			2			0.017			0.014			-0.003			-0.166
1061			Fe			2			12.097			12.849			0.752			0.062
1061			Pb			2			1.84			1.243			-0.597			-0.324
1061			Zn			2			4.754			2.926			-1.828			-0.385
1062			Ag			3			2.318			4.426			2.108			0.909
1062			Au			3			0.023			0.012			-0.012			-0.507
1062			Cu			3			0.162			0.04			-0.121			-0.751
1062			Fe			3			6.15			3.042			-3.107			-0.505
1062			Pb			3			0.402			1.043			0.641			1.597
1062			Zn			3			0.765			2			1.235			1.614
1063			Ag			2			1.651			2.041			0.39			0.236
1063			Au			2			9.18			4.783			-4.397			-0.479
1063			Cu			2			0.01			0.01			0			0.024
1063			Fe			2			17.472			29.482			12.01			0.687
1063			Pb			2			0.012			0.013			0.001			0.114
1063			Zn			2			0.016			0.022			0.007			0.419
1090			Ag			44			76.29			65.609			-10.681			-0.14
1090			Au			44			1.381			1.483			0.101			0.073
1090			Cu			44			0.089			0.073			-0.016			-0.176
1090			Fe			44			16.092			17.873			1.781			0.111
1090			Pb			44			0.697			0.691			-0.005			-0.007
1090			Zn			44			2.463			2.202			-0.261			-0.106
1091			Ag			62			180.145			158.283			-21.862			-0.121
1091			Au			62			0.684			0.694			0.01			0.014

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1091			Cu			62			0.097			0.067			-0.03			-0.313
1091			Fe			62			14.611			14.8			0.189			0.013
1091			Pb			62			1.249			1.116			-0.133			-0.106
1091			Zn			62			2.022			1.789			-0.233			-0.115
1092			Ag			42			27.059			33.562			6.504			0.24
1092			Au			42			1.475			1.477			0.001			0.001
1092			Cu			42			0.088			0.114			0.026			0.29
1092			Fe			42			14.605			15.414			0.809			0.055
1092			Pb			42			0.474			0.442			-0.032			-0.067
1092			Zn			42			1.039			1.016			-0.023			-0.023
1093			Ag			9			71.142			152.319			81.176			1.141
1093			Au			9			1.37			1.806			0.436			0.318
1093			Cu			9			0.095			0.134			0.039			0.405
1093			Fe			9			15.417			16.04			0.624			0.04
1093			Pb			9			0.566			0.507			-0.059			-0.105
1093			Zn			9			1.259			1.605			0.346			0.275
1094			Ag			11			61.518			28.85			-32.668			-0.531
1094			Au			11			1.533			1.829			0.296			0.193
1094			Cu			11			0.059			0.057			-0.001			-0.022
1094			Fe			11			15.294			14.582			-0.712			-0.047
1094			Pb			11			0.855			0.313			-0.542			-0.634
1094			Zn			11			0.526			0.46			-0.066			-0.126
1201			Ag			17			138.451			311.071			172.62			1.247
1201			Au			17			0.913			1.356			0.443			0.485
1201			Cu			17			0.039			0.042			0.003			0.074
1201			Fe			17			14.604			17.245			2.64			0.181
1201			Pb			17			1.437			2.843			1.406			0.979
1201			Zn			17			2.201			2.895			0.694			0.315
1202			Ag			15			44.071			100.007			55.937			1.269
1202			Au			15			0.517			1.352			0.835			1.614
1202			Cu			15			0.115			0.025			-0.09			-0.785
1202			Fe			15			11.9			10.482			-1.418			-0.119
1202			Pb			15			1.009			0.97			-0.039			-0.038
1202			Zn			15			1.56			1.953			0.393			0.252
1203			Ag			15			35.376			58.295			22.919			0.648
1203			Au			15			0.487			1.264			0.777			1.596
1203			Cu			15			0.024			0.028			0.004			0.174
1203			Fe			15			10.587			12.753			2.166			0.205
1203			Pb			15			0.7			0.978			0.277			0.396
1203			Zn			15			1.739			2.279			0.54			0.311
1208			Ag			4			25.537			28.351			2.814			0.11
1208			Au			4			0.863			0.768			-0.095			-0.11
1208			Cu			4			0.086			0.057			-0.029			-0.34
1208			Fe			4			16.077			16.362			0.285			0.018
1208			Pb			4			0.252			0.186			-0.066			-0.261
1208			Zn			4			0.239			0.174			-0.065			-0.272
1209			Ag			6			15.177			16.39			1.213			0.08
1209			Au			6			0.303			0.44			0.137			0.451
1209			Cu			6			0.173			0.117			-0.056			-0.323
1209			Fe			6			6.489			5.943			-0.546			-0.084
1209			Pb			6			0.27			0.278			0.008			0.029

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        468

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1209			Zn			6			1.143			1.076			-0.067			-0.059
1304			Ag			6			18.882			14.93			-3.952			-0.209
1304			Au			6			1.224			0.547			-0.677			-0.553
1304			Cu			6			0.055			0.034			-0.021			-0.377
1304			Fe			6			14.602			12.578			-2.024			-0.139
1304			Pb			6			0.634			0.762			0.129			0.203
1304			Zn			6			0.74			0.685			-0.054			-0.073
1305			Ag			29			25.308			26.546			1.238			0.049
1305			Au			29			1.142			0.997			-0.145			-0.127
1305			Cu			29			0.085			0.069			-0.016			-0.191
1305			Fe			29			18.845			17.485			-1.36			-0.072
1305			Pb			29			0.276			0.24			-0.036			-0.132
1305			Zn			29			0.758			0.493			-0.265			-0.35
1306			Ag			10			50.538			69.685			19.147			0.379
1306			Au			10			0.57			0.841			0.271			0.475
1306			Cu			10			0.057			0.071			0.014			0.242
1306			Fe			10			10.368			11.145			0.777			0.075
1306			Pb			10			0.574			0.342			-0.231			-0.403
1306			Zn			10			0.829			0.615			-0.213			-0.257
1310			Ag			18			23.585			18.649			-4.936			-0.209
1310			Au			18			1.045			1.118			0.073			0.07
1310			Cu			18			0.073			0.075			0.001			0.019
1310			Fe			18			14.076			17.07			2.994			0.213
1310			Pb			18			0.494			0.311			-0.183			-0.37
1310			Zn			18			1.86			0.68			-1.18			-0.634
1311			Ag			6			26.335			33.47			7.135			0.271
1311			Au			6			0.568			1.1			0.532			0.938
1311			Cu			6			0.068			0.079			0.01			0.149
1311			Fe			6			20.791			23.431			2.64			0.127
1311			Pb			6			0.117			0.083			-0.034			-0.292
1311			Zn			6			0.423			0.319			-0.104			-0.245
1312			Ag			13			22.733			15.166			-7.567			-0.333
1312			Au			13			0.349			0.41			0.061			0.174
1312			Cu			13			0.043			0.049			0.006			0.139
1312			Fe			13			7.715			6.399			-1.316			-0.171
1312			Pb			13			0.365			0.243			-0.123			-0.336
1312			Zn			13			1.086			1.458			0.373			0.343
1313			Ag			8			12.043			21.926			9.883			0.821
1313			Au			8			0.156			0.139			-0.017			-0.112
1313			Cu			8			0.027			0.023			-0.005			-0.167
1313			Fe			8			7.921			6.154			-1.767			-0.223
1313			Pb			8			0.766			1.571			0.805			1.051
1313			Zn			8			1.513			1.29			-0.223			-0.147
1314			Ag			8			13.727			13.457			-0.27			-0.02
1314			Au			8			0.714			0.712			-0.002			-0.003
1314			Cu			8			0.068			0.071			0.003			0.04
1314			Fe			8			14.609			14.235			-0.374			-0.026
1314			Pb			8			0.182			0.181			-0.001			-0.006
1314			Zn			8			0.288			0.287			-0.001			-0.003
1331			Ag			14			28.39			27.687			-0.702			-0.025
1331			Au			14			0.64			1.206			0.567			0.886

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        469

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1331			Cu			14			0.055			0.079			0.024			0.445
1331			Fe			14			12.13			15.118			2.987			0.246
1331			Pb			14			1.099			0.977			-0.121			-0.111
1331			Zn			14			1.957			1.292			-0.665			-0.34
1341			Ag			14			13.475			17.724			4.249			0.315
1341			Au			14			0.319			0.299			-0.02			-0.062
1341			Cu			14			0.043			0.059			0.015			0.358
1341			Fe			14			8.902			7.798			-1.104			-0.124
1341			Pb			14			0.765			0.991			0.226			0.296
1341			Zn			14			1.485			1.487			0.002			0.001
1403			Ag			7			18.74			18.585			-0.155			-0.008
1403			Au			7			0.311			0.392			0.081			0.26
1403			Cu			7			0.046			0.044			-0.002			-0.037
1403			Fe			7			17.885			16.088			-1.797			-0.1
1403			Pb			7			1.651			1.931			0.281			0.17
1403			Zn			7			2.066			1.684			-0.382			-0.185
1404			Ag			15			35.756			29.244			-6.513			-0.182
1404			Au			15			0.349			0.197			-0.152			-0.436
1404			Cu			15			0.125			0.096			-0.029			-0.234
1404			Fe			15			15.532			15.534			0.002			0
1404			Pb			15			0.572			0.439			-0.133			-0.232
1404			Zn			15			2.113			1.234			-0.879			-0.416
1407			Ag			17			17.809			21.77			3.961			0.222
1407			Au			17			0.639			1.2			0.56			0.876
1407			Cu			17			0.059			0.086			0.027			0.456
1407			Fe			17			14.893			18.028			3.135			0.211
1407			Pb			17			0.458			0.429			-0.029			-0.063
1407			Zn			17			0.733			0.83			0.096			0.132
1501			Ag			12			70.94			72.539			1.598			0.023
1501			Au			12			0.823			1.129			0.307			0.373
1501			Cu			12			0.032			0.056			0.024			0.765
1501			Fe			12			16.317			16.132			-0.185			-0.011
1501			Pb			12			0.897			0.81			-0.087			-0.097
1501			Zn			12			1.68			3.153			1.473			0.877
1502			Ag			19			100.706			126.634			25.928			0.257
1502			Au			19			0.582			0.708			0.126			0.217
1502			Cu			19			0.033			0.032			-0.002			-0.048
1502			Fe			19			12.732			13.707			0.975			0.077
1502			Pb			19			0.987			1.253			0.265			0.269
1502			Zn			19			2.468			3.237			0.769			0.312
1503			Ag			18			79.693			62.083			-17.61			-0.221
1503			Au			18			0.518			0.612			0.094			0.181
1503			Cu			18			0.054			0.04			-0.014			-0.259
1503			Fe			18			14.953			15.957			1.003			0.067
1503			Pb			18			0.738			0.625			-0.113			-0.153
1503			Zn			18			1.711			1.465			-0.246			-0.144
1504			Ag			8			119.388			119.237			-0.151			-0.001
1504			Au			8			0.802			1.179			0.376			0.469
1504			Cu			8			0.027			0.023			-0.004			-0.135
1504			Fe			8			15.004			14.011			-0.993			-0.066
1504			Pb			8			0.914			1.103			0.189			0.207

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        470

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1504			Zn			8			1.555			1.71			0.154			0.099
1505			Ag			10			69.453			70.657			1.204			0.017
1505			Au			10			0.599			0.592			-0.007			-0.011
1505			Cu			10			0.03			0.029			-0.001			-0.047
1505			Fe			10			14.318			15.654			1.336			0.093
1505			Pb			10			0.49			0.714			0.223			0.455
1505			Zn			10			1.156			0.946			-0.21			-0.182
1601			Ag			6			37.935			22.985			-14.95			-0.394
1601			Au			6			0.234			0.203			-0.031			-0.133
1601			Cu			6			0.149			0.092			-0.057			-0.382
1601			Fe			6			16.423			16.363			-0.061			-0.004
1601			Pb			6			2.295			1.235			-1.06			-0.462
1601			Zn			6			2.595			1.936			-0.659			-0.254
1614			Ag			5			23.656			21.429			-2.227			-0.094
1614			Au			5			0.625			0.619			-0.006			-0.01
1614			Cu			5			0.115			0.065			-0.05			-0.437
1614			Fe			5			16.995			16.937			-0.057			-0.003
1614			Pb			5			0.135			0.128			-0.008			-0.056
1614			Zn			5			0.15			0.122			-0.028			-0.186
1631			Ag			6			25.869			27.727			1.858			0.072
1631			Au			6			0.518			0.648			0.13			0.251
1631			Cu			6			0.034			0.018			-0.016			-0.476
1631			Fe			6			11.692			10.96			-0.732			-0.063
1631			Pb			6			0.643			0.443			-0.199			-0.31
1631			Zn			6			0.686			0.594			-0.091			-0.133
1632			Ag			15			19.574			20.809			1.235			0.063
1632			Au			15			0.943			0.977			0.035			0.037
1632			Cu			15			0.033			0.039			0.006			0.187
1632			Fe			15			11.956			10.91			-1.046			-0.087
1632			Pb			15			0.556			0.269			-0.288			-0.517
1632			Zn			15			1.392			0.559			-0.834			-0.599
2002			Ag			21			26.508			23.63			-2.879			-0.109
2002			Au			21			1.143			2.025			0.882			0.771
2002			Cu			21			0.046			0.061			0.016			0.341
2002			Fe			21			14.844			13.801			-1.042			-0.07
2002			Pb			21			0.924			0.514			-0.411			-0.444
2002			Zn			21			1.101			0.918			-0.183			-0.166
2004			Ag			9			67.003			73.827			6.824			0.102
2004			Au			9			1.135			1.455			0.32			0.282
2004			Cu			9			0.147			0.191			0.044			0.3
2004			Fe			9			17.733			17.992			0.259			0.015
2004			Pb			9			0.328			0.383			0.055			0.168
2004			Zn			9			1.839			1.611			-0.228			-0.124
2005			Ag			11			23.936			17.389			-6.547			-0.274
2005			Au			11			0.219			0.32			0.101			0.461
2005			Cu			11			0.014			0.014			0			-0.001
2005			Fe			11			14.874			14.324			-0.55			-0.037
2005			Pb			11			0.405			0.477			0.073			0.18
2005			Zn			11			1.488			1.831			0.343			0.231
2006			Ag			9			36.885			16.973			-19.912			-0.54
2006			Au			9			0.282			0.846			0.564			2

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        471

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2006			Cu			9			0.029			0.026			-0.003			-0.1
2006			Fe			9			10.756			12.808			2.052			0.191
2006			Pb			9			1.824			0.963			-0.862			-0.472
2006			Zn			9			1.634			1.438			-0.196			-0.12
2009			Ag			14			40.554			25.07			-15.485			-0.382
2009			Au			14			1.561			0.905			-0.656			-0.42
2009			Cu			14			0.059			0.036			-0.023			-0.397
2009			Fe			14			17.293			15.63			-1.664			-0.096
2009			Pb			14			0.901			0.505			-0.397			-0.44
2009			Zn			14			1.181			0.89			-0.29			-0.246
2012			Ag			13			37.485			37.831			0.345			0.009
2012			Au			13			0.584			0.775			0.191			0.328
2012			Cu			13			0.025			0.036			0.011			0.449
2012			Fe			13			13.047			12.846			-0.201			-0.015
2012			Pb			13			1.116			1.037			-0.078			-0.07
2012			Zn			13			2.778			3.416			0.638			0.23
2018			Ag			9			57.536			79.426			21.89			0.38
2018			Au			9			1.486			2.408			0.922			0.62
2018			Cu			9			0.03			0.043			0.013			0.419
2018			Fe			9			17.03			25.585			8.555			0.502
2018			Pb			9			0.744			0.613			-0.131			-0.176
2018			Zn			9			3.691			2.868			-0.823			-0.223
2019			Ag			4			78.913			58.768			-20.145			-0.255
2019			Au			4			1.697			2.029			0.332			0.195
2019			Cu			4			0.161			0.147			-0.014			-0.088
2019			Fe			4			16.958			22.49			5.532			0.326
2019			Pb			4			0.238			0.203			-0.035			-0.148
2019			Zn			4			0.46			0.581			0.121			0.263
2020			Ag			8			30.084			46.517			16.433			0.546
2020			Au			8			2.327			4.203			1.877			0.807
2020			Cu			8			0.045			0.064			0.018			0.405
2020			Fe			8			14.912			17.952			3.04			0.204
2020			Pb			8			0.342			0.376			0.034			0.098
2020			Zn			8			1.115			0.72			-0.395			-0.354
2026			Ag			4			14.069			13.826			-0.243			-0.017
2026			Au			4			0.342			0.534			0.192			0.562
2026			Cu			4			0.013			0.016			0.003			0.246
2026			Fe			4			12.246			13.743			1.497			0.122
2026			Pb			4			0.347			0.794			0.447			1.288
2026			Zn			4			1.052			1.912			0.86			0.818
2091			Ag			118			53.233			45.397			-7.836			-0.147
2091			Au			118			2.853			2.455			-0.399			-0.14
2091			Cu			118			0.085			0.082			-0.003			-0.037
2091			Fe			118			17.978			17.516			-0.462			-0.026
2091			Pb			118			0.593			0.522			-0.071			-0.12
2091			Zn			118			1.722			1.542			-0.18			-0.105
2101			Ag			8			41.468			38.322			-3.145			-0.076
2101			Au			8			1.363			1.221			-0.142			-0.104
2101			Cu			8			0.048			0.044			-0.004			-0.078
2101			Fe			8			12.145			11.766			-0.379			-0.031
2101			Pb			8			0.435			0.46			0.024			0.056

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        472

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2101			Zn			8			2.401			2.821			0.42			0.175
2103			Ag			32			39.127			35.979			-3.148			-0.08
2103			Au			32			0.806			0.952			0.146			0.181
2103			Cu			32			0.061			0.051			-0.011			-0.171
2103			Fe			32			14.083			16.109			2.026			0.144
2103			Pb			32			0.847			0.57			-0.277			-0.327
2103			Zn			32			2.401			1.711			-0.689			-0.287
2104			Ag			39			38.454			49.081			10.627			0.276
2104			Au			39			1.216			1.769			0.553			0.455
2104			Cu			39			0.085			0.09			0.005			0.058
2104			Fe			39			14.677			19.514			4.837			0.33
2104			Pb			39			0.646			0.553			-0.093			-0.144
2104			Zn			39			2.018			1.809			-0.209			-0.104
2106			Ag			18			24.79			16.758			-8.032			-0.324
2106			Au			18			1.784			1.058			-0.725			-0.407
2106			Cu			18			0.106			0.048			-0.058			-0.549
2106			Fe			18			14.966			11.653			-3.313			-0.221
2106			Pb			18			0.457			0.303			-0.153			-0.335
2106			Zn			18			1.35			0.916			-0.434			-0.321
2113			Ag			6			27.045			23.665			-3.38			-0.125
2113			Au			6			0.756			0.606			-0.15			-0.199
2113			Cu			6			0.064			0.053			-0.011			-0.167
2113			Fe			6			13.277			11.879			-1.399			-0.105
2113			Pb			6			0.758			0.636			-0.122			-0.161
2113			Zn			6			1.081			1.084			0.003			0.003
2115			Ag			6			17			18.138			1.138			0.067
2115			Au			6			0.275			0.26			-0.015			-0.054
2115			Cu			6			0.034			0.035			0.002			0.048
2115			Fe			6			12.384			13.413			1.03			0.083
2115			Pb			6			0.523			0.599			0.077			0.146
2115			Zn			6			1.208			1.307			0.099			0.082
2117			Ag			23			58.768			71.693			12.925			0.22
2117			Au			23			1.806			2.485			0.679			0.376
2117			Cu			23			0.098			0.133			0.035			0.358
2117			Fe			23			16.809			17.398			0.588			0.035
2117			Pb			23			0.71			0.483			-0.227			-0.319
2117			Zn			23			1.546			0.771			-0.775			-0.501
2118			Ag			8			25.872			28.667			2.795			0.108
2118			Au			8			0.803			1.2			0.397			0.494
2118			Cu			8			0.059			0.091			0.032			0.535
2118			Fe			8			14.142			22.868			8.725			0.617
2118			Pb			8			0.596			0.464			-0.132			-0.221
2118			Zn			8			1.921			1.123			-0.798			-0.415
2120			Ag			7			89.449			77.787			-11.662			-0.13
2120			Au			7			2.524			2.787			0.263			0.104
2120			Cu			7			0.2			0.169			-0.031			-0.155
2120			Fe			7			20.178			21.365			1.186			0.059
2120			Pb			7			0.462			0.546			0.083			0.18
2120			Zn			7			1.175			1.677			0.502			0.427
2122			Ag			6			208.21			65.847			-142.363			-0.684
2122			Au			6			0.796			1.431			0.635			0.797

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        473

---

2122			Cu			6			0.069			0.054			-0.015			-0.218
2122			Fe			6			14.47			21.296			6.826			0.472
2122			Pb			6			2.192			1.106			-1.086			-0.495
2122			Zn			6			0.446			0.627			0.181			0.405
2127			Ag			2			39.5			22.478			-17.022			-0.431
2127			Au			2			0.135			0.075			-0.06			-0.448
2127			Cu			2			0.21			0.119			-0.091			-0.435
2127			Fe			2			16.726			10.638			-6.089			-0.364
2127			Pb			2			0.129			0.153			0.024			0.186
2127			Zn			2			1.157			1.681			0.524			0.453
2130			Ag			7			24.558			15.776			-8.782			-0.358
2130			Au			7			1.558			0.739			-0.819			-0.526
2130			Cu			7			0.097			0.147			0.05			0.516
2130			Fe			7			15.189			9.078			-6.11			-0.402
2130			Pb			7			1.478			1.25			-0.228			-0.154
2130			Zn			7			1.687			2.302			0.615			0.365
2131			Ag			6			19.957			17.587			-2.371			-0.119
2131			Au			6			2.131			1.746			-0.385			-0.181
2131			Cu			6			0.027			0.029			0.001			0.054
2131			Fe			6			20.165			22.919			2.754			0.137
2131			Pb			6			0.483			0.48			-0.003			-0.006
2131			Zn			6			1.123			1.554			0.431			0.384
2132			Ag			12			18.931			25.137			6.206			0.328
2132			Au			12			0.417			0.64			0.223			0.536
2132			Cu			12			0.039			0.065			0.026			0.654
2132			Fe			12			9.943			12.086			2.143			0.216
2132			Pb			12			0.616			0.65			0.034			0.055
2132			Zn			12			1.177			1.058			-0.118			-0.101
2203			Ag			10			96.192			46.891			-49.301			-0.513
2203			Au			10			0.404			0.363			-0.042			-0.103
2203			Cu			10			0.013			0.012			-0.001			-0.08
2203			Fe			10			12.686			13.154			0.468			0.037
2203			Pb			10			1.398			0.878			-0.521			-0.372
2203			Zn			10			3.171			3.04			-0.131			-0.041
2207			Ag			5			51.618			47.062			-4.556			-0.088
2207			Au			5			1.233			1.311			0.078			0.063
2207			Cu			5			0.01			0.01			0			0.024
2207			Fe			5			12.226			7.956			-4.271			-0.349
2207			Pb			5			0.467			0.422			-0.045			-0.097
2207			Zn			5			1.177			0.486			-0.692			-0.587
2213			Ag			4			37.305			35.18			-2.125			-0.057
2213			Au			4			2.405			1.973			-0.432			-0.18
2213			Cu			4			0.104			0.055			-0.049			-0.474
2213			Fe			4			15.064			10.333			-4.731			-0.314
2213			Pb			4			0.359			0.32			-0.039			-0.109
2213			Zn			4			0.482			0.697			0.215			0.446
2215			Ag			4			4.797			5.602			0.805			0.168
2215			Au			4			0.159			0.095			-0.064			-0.401
2215			Cu			4			0.012			0.013			0.001			0.073
2215			Fe			4			12.417			9.901			-2.516			-0.203
2215			Pb			4			0.243			0.276			0.034			0.138

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        474

---

2215			Zn			4			2.904			4.02			1.116			0.384
2216			Ag			10			21.086			20.471			-0.615			-0.029
2216			Au			10			0.717			0.477			-0.239			-0.334
2216			Cu			10			0.013			0.014			0.001			0.055
2216			Fe			10			10.24			9.907			-0.333			-0.033
2216			Pb			10			0.43			0.333			-0.097			-0.226
2216			Zn			10			1.148			1.505			0.357			0.311
2219			Ag			9			68.299			98.258			29.959			0.439
2219			Au			9			2.123			3.665			1.542			0.726
2219			Cu			9			0.121			0.172			0.052			0.429
2219			Fe			9			22.521			31.386			8.865			0.394
2219			Pb			9			1.848			0.988			-0.86			-0.465
2219			Zn			9			0.552			1.036			0.484			0.877
2221			Ag			7			24.914			26.571			1.657			0.066
2221			Au			7			0.488			0.558			0.069			0.142
2221			Cu			7			0.022			0.024			0.002			0.087
2221			Fe			7			11.272			11.021			-0.251			-0.022
2221			Pb			7			0.798			0.854			0.056			0.07
2221			Zn			7			0.868			0.787			-0.082			-0.094
2222			Ag			5			15.722			15.854			0.132			0.008
2222			Au			5			0.243			0.12			-0.123			-0.506
2222			Cu			5			0.05			0.063			0.013			0.267
2222			Fe			5			9.672			4.298			-5.374			-0.556
2222			Pb			5			0.531			0.841			0.31			0.585
2222			Zn			5			5.083			9.063			3.98			0.783
2223			Ag			8			81.908			31.693			-50.214			-0.613
2223			Au			8			0.941			0.982			0.041			0.044
2223			Cu			8			0.024			0.016			-0.009			-0.358
2223			Fe			8			10.776			11.975			1.199			0.111
2223			Pb			8			0.725			0.611			-0.114			-0.157
2223			Zn			8			1.44			1.617			0.177			0.123
2224			Ag			2			35.93			30.328			-5.602			-0.156
2224			Au			2			0.6			0.941			0.341			0.569
2224			Cu			2			0.036			0.023			-0.014			-0.372
2224			Fe			2			6.92			7.947			1.026			0.148
2224			Pb			2			3.477			1.605			-1.872			-0.538
2224			Zn			2			0.621			0.819			0.198			0.319
2225			Ag			6			27.299			33.416			6.117			0.224
2225			Au			6			2.714			2.907			0.193			0.071
2225			Cu			6			0.062			0.077			0.015			0.246
2225			Fe			6			18.796			21.567			2.771			0.147
2225			Pb			6			0.254			0.294			0.04			0.156
2225			Zn			6			0.713			0.421			-0.292			-0.41
2235			Ag			5			11.755			14.143			2.389			0.203
2235			Au			5			3.347			3.273			-0.074			-0.022
2235			Cu			5			0.024			0.024			-0.001			-0.029
2235			Fe			5			8.832			10.162			1.329			0.151
2235			Pb			5			0.432			0.439			0.006			0.014
2235			Zn			5			2.226			2.41			0.184			0.083
2236			Ag			2			52.923			61.065			8.142			0.154
2236			Au			2			0.476			0.593			0.117			0.246

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        475

---

2236			Cu			2			0.02			0.025			0.005			0.245
2236			Fe			2			16.528			15.226			-1.302			-0.079
2236			Pb			2			0.448			0.485			0.037			0.083
2236			Zn			2			1.137			1.316			0.179			0.157
2309			Ag			7			78.947			63.33			-15.618			-0.198
2309			Au			7			1.243			1.014			-0.229			-0.184
2309			Cu			7			0.09			0.052			-0.038			-0.419
2309			Fe			7			19.435			16.026			-3.41			-0.175
2309			Pb			7			1.44			0.53			-0.91			-0.632
2309			Zn			7			0.694			0.3			-0.394			-0.568
2311			Ag			13			28.545			26.104			-2.441			-0.086
2311			Au			13			1.187			1.11			-0.078			-0.065
2311			Cu			13			0.044			0.043			-0.001			-0.026
2311			Fe			13			15.921			13.932			-1.989			-0.125
2311			Pb			13			0.593			0.494			-0.099			-0.167
2311			Zn			13			1.71			1.368			-0.342			-0.2
2314			Ag			4			73.161			57.317			-15.844			-0.217
2314			Au			4			0.603			0.598			-0.005			-0.009
2314			Cu			4			0.142			0.138			-0.004			-0.029
2314			Fe			4			22.157			25.913			3.756			0.17
2314			Pb			4			0.887			0.882			-0.005			-0.005
2314			Zn			4			1.745			2.509			0.764			0.438
2327			Ag			6			39.443			35.109			-4.334			-0.11
2327			Au			6			0.826			0.914			0.088			0.107
2327			Cu			6			0.019			0.019			-0.001			-0.033
2327			Fe			6			15.266			17.324			2.058			0.135
2327			Pb			6			0.312			0.25			-0.062			-0.2
2327			Zn			6			1.766			1.178			-0.588			-0.333
2400			Ag			20			68.845			62.483			-6.362			-0.092
2400			Au			20			1.191			1.15			-0.041			-0.035
2400			Cu			20			0.065			0.052			-0.013			-0.196
2400			Fe			20			17.745			17.228			-0.517			-0.029
2400			Pb			20			0.56			0.465			-0.095			-0.17
2400			Zn			20			0.62			0.754			0.134			0.217
2401			Ag			18			36.651			43.911			7.259			0.198
2401			Au			18			0.252			0.404			0.152			0.604
2401			Cu			18			0.029			0.042			0.013			0.465
2401			Fe			18			11.424			10.957			-0.467			-0.041
2401			Pb			18			1.776			1.812			0.035			0.02
2401			Zn			18			1.129			0.978			-0.15			-0.133
2405			Ag			5			28.589			21.057			-7.532			-0.263
2405			Au			5			0.316			0.202			-0.114			-0.36
2405			Cu			5			0.02			0.026			0.006			0.31
2405			Fe			5			15.555			18.318			2.763			0.178
2405			Pb			5			0.552			0.792			0.24			0.435
2405			Zn			5			0.637			0.806			0.169			0.265
2408			Ag			7			52.679			50.904			-1.774			-0.034
2408			Au			7			1.203			0.872			-0.332			-0.275
2408			Cu			7			0.053			0.057			0.004			0.067
2408			Fe			7			13.325			10.782			-2.543			-0.191
2408			Pb			7			0.323			0.342			0.019			0.059

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        476

---

2408			Zn			7			0.508			0.517			0.009			0.018
2410			Ag			13			43.166			80.849			37.684			0.873
2410			Au			13			0.757			1.707			0.951			1.256
2410			Cu			13			0.028			0.056			0.027			0.959
2410			Fe			13			15.677			18.964			3.286			0.21
2410			Pb			13			0.8			0.837			0.038			0.047
2410			Zn			13			2.003			1.86			-0.143			-0.071
2411			Ag			6			22.431			19.751			-2.68			-0.119
2411			Au			6			0.903			0.982			0.079			0.087
2411			Cu			6			0.046			0.036			-0.01			-0.213
2411			Fe			6			20.349			20.551			0.202			0.01
2411			Pb			6			0.446			0.342			-0.104			-0.233
2411			Zn			6			0.863			0.757			-0.106			-0.123
2417			Ag			4			44.163			30.685			-13.478			-0.305
2417			Au			4			1.81			1.211			-0.599			-0.331
2417			Cu			4			0.015			0.012			-0.002			-0.144
2417			Fe			4			19.297			19.69			0.393			0.02
2417			Pb			4			0.448			0.268			-0.181			-0.403
2417			Zn			4			0.789			0.47			-0.319			-0.404
2424			Ag			4			27.333			22.051			-5.282			-0.193
2424			Au			4			0.594			0.614			0.02			0.034
2424			Cu			4			0.018			0.019			0.001			0.042
2424			Fe			4			17.448			15.106			-2.343			-0.134
2424			Pb			4			0.701			0.592			-0.109			-0.156
2424			Zn			4			1.742			1.424			-0.318			-0.183
2425			Ag			4			39.579			43.547			3.968			0.1
2425			Au			4			0.263			0.398			0.135			0.516
2425			Cu			4			0.315			0.191			-0.125			-0.396
2425			Fe			4			2.369			2.025			-0.345			-0.145
2425			Pb			4			2.58			2.435			-0.145			-0.056
2425			Zn			4			0.955			0.969			0.013			0.014
2502			Ag			11			28.064			21.315			-6.748			-0.24
2502			Au			11			0.698			0.839			0.141			0.202
2502			Cu			11			0.248			0.068			-0.18			-0.724
2502			Fe			11			15.97			16.075			0.105			0.007
2502			Pb			11			1.507			0.705			-0.801			-0.532
2502			Zn			11			1.32			0.84			-0.481			-0.364
2507			Ag			23			52.038			66.802			14.764			0.284
2507			Au			23			1.087			2.424			1.337			1.23
2507			Cu			23			0.105			0.089			-0.016			-0.152
2507			Fe			23			14.921			17.944			3.023			0.203
2507			Pb			23			0.739			0.724			-0.014			-0.019
2507			Zn			23			1.784			1.821			0.037			0.021
2510			Ag			7			21.334			21.873			0.539			0.025
2510			Au			7			0.521			0.649			0.127			0.245
2510			Cu			7			0.074			0.116			0.043			0.58
2510			Fe			7			17.277			20.883			3.606			0.209
2510			Pb			7			0.348			0.351			0.003			0.01
2510			Zn			7			0.399			0.411			0.012			0.03
2538			Ag			9			20.913			14.832			-6.081			-0.291
2538			Au			9			0.188			0.161			-0.026			-0.139

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        477

---

2538			Cu			9			0.02			0.027			0.007			0.328
2538			Fe			9			11.743			7.469			-4.274			-0.364
2538			Pb			9			0.804			1.001			0.197			0.245
2538			Zn			9			2.478			3.507			1.028			0.415
2614			Ag			4			30.864			23.221			-7.643			-0.248
2614			Au			4			0.215			0.215			0			-0.002
2614			Cu			4			0.07			0.16			0.09			1.287
2614			Fe			4			20.711			16.634			-4.077			-0.197
2614			Pb			4			0.483			0.487			0.004			0.009
2614			Zn			4			1.352			2.018			0.667			0.493
2729			Ag			4			24.075			26.631			2.556			0.106
2729			Au			4			1.26			1.412			0.152			0.12
2729			Cu			4			0.07			0.077			0.008			0.109
2729			Fe			4			20.68			23.654			2.974			0.144
2729			Pb			4			0.482			0.361			-0.121			-0.252
2729			Zn			4			0.472			0.442			-0.03			-0.063
2802			Ag			6			39.183			44.573			5.39			0.138
2802			Au			6			0.698			1.038			0.34			0.486
2802			Cu			6			0.039			0.035			-0.004			-0.093
2802			Fe			6			11.077			12.646			1.57			0.142
2802			Pb			6			1.989			1.679			-0.31			-0.156
2802			Zn			6			2.506			3.661			1.156			0.461
3001			Ag			27			19.677			29.821			10.144			0.516
3001			Au			27			0.618			0.628			0.01			0.016
3001			Cu			27			0.037			0.049			0.013			0.344
3001			Fe			27			13.147			15.181			2.034			0.155
3001			Pb			27			0.408			0.227			-0.182			-0.445
3001			Zn			27			0.831			0.526			-0.305			-0.367
3002			Ag			7			28.225			25.017			-3.208			-0.114
3002			Au			7			0.809			0.704			-0.105			-0.13
3002			Cu			7			0.028			0.023			-0.005			-0.165
3002			Fe			7			13.037			13.059			0.022			0.002
3002			Pb			7			0.619			0.82			0.201			0.325
3002			Zn			7			1.444			1.378			-0.066			-0.046
3003			Ag			8			12.122			11.834			-0.288			-0.024
3003			Au			8			0.175			0.148			-0.027			-0.152
3003			Cu			8			0.025			0.027			0.002			0.1
3003			Fe			8			10.897			11.422			0.524			0.048
3003			Pb			8			0.26			0.252			-0.008			-0.03
3003			Zn			8			0.477			0.515			0.037			0.079
3004			Ag			7			16.008			19.307			3.299			0.206
3004			Au			7			0.531			0.757			0.226			0.426
3004			Cu			7			0.021			0.03			0.009			0.446
3004			Fe			7			10.042			10.048			0.006			0.001
3004			Pb			7			0.301			0.458			0.157			0.522
3004			Zn			7			0.569			0.746			0.177			0.312
3005			Ag			11			23.019			22.089			-0.931			-0.04
3005			Au			11			0.94			1.117			0.177			0.188
3005			Cu			11			0.022			0.018			-0.004			-0.161
3005			Fe			11			13.15			14.629			1.479			0.113
3005			Pb			11			0.347			0.295			-0.052			-0.15

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        478

---

3005			Zn			11			0.777			0.766			-0.011			-0.014
3006			Ag			10			21.284			32.247			10.963			0.515
3006			Au			10			0.5			0.521			0.021			0.043
3006			Cu			10			0.078			0.077			0			-0.006
3006			Fe			10			15.81			22.135			6.325			0.4
3006			Pb			10			0.538			0.38			-0.158			-0.294
3006			Zn			10			2.279			2.778			0.499			0.219
3007			Ag			7			24.601			27.105			2.504			0.102
3007			Au			7			0.209			0.257			0.048			0.227
3007			Cu			7			0.06			0.062			0.003			0.045
3007			Fe			7			14.749			13.065			-1.684			-0.114
3007			Pb			7			0.357			0.465			0.108			0.304
3007			Zn			7			0.944			0.976			0.033			0.035
3008			Ag			5			46.08			40.136			-5.944			-0.129
3008			Au			5			0.222			0.144			-0.078			-0.352
3008			Cu			5			0.02			0.018			-0.002			-0.114
3008			Fe			5			17.965			20.246			2.281			0.127
3008			Pb			5			0.385			0.259			-0.127			-0.329
3008			Zn			5			1.217			0.91			-0.307			-0.252
3009			Ag			13			20.921			27.766			6.845			0.327
3009			Au			13			0.824			1.911			1.087			1.319
3009			Cu			13			0.024			0.037			0.013			0.565
3009			Fe			13			15.947			16.938			0.99			0.062
3009			Pb			13			0.378			0.388			0.011			0.028
3009			Zn			13			0.573			0.507			-0.066			-0.116
3010			Ag			12			16.781			18.476			1.695			0.101
3010			Au			12			1.219			1.296			0.077			0.063
3010			Cu			12			0.043			0.053			0.011			0.25
3010			Fe			12			11.118			10.684			-0.434			-0.039
3010			Pb			12			0.38			0.291			-0.089			-0.233
3010			Zn			12			0.539			0.395			-0.145			-0.268
3011			Ag			8			15.821			17.584			1.763			0.111
3011			Au			8			0.588			0.565			-0.023			-0.039
3011			Cu			8			0.027			0.032			0.004			0.163
3011			Fe			8			12.779			12.353			-0.426			-0.033
3011			Pb			8			0.435			0.476			0.041			0.095
3011			Zn			8			1.774			1.947			0.173			0.098
3100			Ag			2			16.833			19.072			2.239			0.133
3100			Au			2			0.202			0.154			-0.048			-0.238
3100			Cu			2			0.011			0.01			0			-0.022
3100			Fe			2			6.755			6.854			0.099			0.015
3100			Pb			2			0.596			0.7			0.104			0.175
3100			Zn			2			3.476			4.742			1.266			0.364
3101			Ag			46			30.546			30.74			0.194			0.006
3101			Au			46			1.109			1.205			0.096			0.086
3101			Cu			46			0.056			0.072			0.016			0.278
3101			Fe			46			15.401			17.084			1.683			0.109
3101			Pb			46			0.325			0.291			-0.034			-0.103
3101			Zn			46			0.913			0.777			-0.137			-0.149
3102			Ag			17			37.569			49.928			12.359			0.329
3102			Au			17			0.816			0.869			0.053			0.065

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        479

---

3102			Cu			17			0.042			0.049			0.007			0.169
3102			Fe			17			14.514			16.471			1.957			0.135
3102			Pb			17			0.432			0.357			-0.075			-0.174
3102			Zn			17			0.811			0.701			-0.109			-0.135
3103			Ag			21			22.738			30.128			7.389			0.325
3103			Au			21			1.313			1.388			0.075			0.057
3103			Cu			21			0.047			0.101			0.055			1.171
3103			Fe			21			16.856			20.094			3.238			0.192
3103			Pb			21			0.255			0.173			-0.082			-0.321
3103			Zn			21			0.618			0.409			-0.208			-0.337
3121			Ag			20			34.323			26.172			-8.151			-0.237
3121			Au			20			1.182			1.369			0.187			0.158
3121			Cu			20			0.119			0.112			-0.006			-0.051
3121			Fe			20			19.549			21.369			1.82			0.093
3121			Pb			20			0.48			0.308			-0.172			-0.358
3121			Zn			20			1.018			0.823			-0.194			-0.191
3191			Ag			77			41.141			35.812			-5.329			-0.13
3191			Au			77			1.713			1.714			0.001			0.001
3191			Cu			77			0.084			0.07			-0.014			-0.162
3191			Fe			77			16.96			16.823			-0.137			-0.008
3191			Pb			77			0.518			0.318			-0.2			-0.386
3191			Zn			77			1.206			0.879			-0.328			-0.271
3192			Ag			17			53.341			29.819			-23.522			-0.441
3192			Au			17			1.541			1.213			-0.328			-0.213
3192			Cu			17			0.072			0.047			-0.025			-0.344
3192			Fe			17			22.151			23.727			1.576			0.071
3192			Pb			17			0.375			0.153			-0.222			-0.592
3192			Zn			17			1.218			0.36			-0.859			-0.705
3193			Ag			16			23.467			20.547			-2.92			-0.124
3193			Au			16			1.202			0.869			-0.333			-0.277
3193			Cu			16			0.034			0.044			0.009			0.269
3193			Fe			16			16.45			25.233			8.783			0.534
3193			Pb			16			0.369			0.285			-0.084			-0.227
3193			Zn			16			1.061			1.512			0.451			0.425
3204			Ag			15			46.945			28.169			-18.776			-0.4
3204			Au			15			1.035			1.119			0.084			0.081
3204			Cu			15			0.083			0.039			-0.044			-0.531
3204			Fe			15			16.559			16.659			0.099			0.006
3204			Pb			15			0.351			0.219			-0.132			-0.375
3204			Zn			15			0.723			0.932			0.209			0.289
3205			Ag			15			21.495			31.025			9.53			0.443
3205			Au			15			0.343			0.224			-0.12			-0.348
3205			Cu			15			0.029			0.026			-0.003			-0.101
3205			Fe			15			16.537			15.825			-0.712			-0.043
3205			Pb			15			0.411			0.363			-0.048			-0.117
3205			Zn			15			1.311			1.138			-0.173			-0.132
3206			Ag			18			27.744			16.404			-11.34			-0.409
3206			Au			18			0.624			0.696			0.071			0.114
3206			Cu			18			0.049			0.05			0			0.008
3206			Fe			18			18.044			14.377			-3.667			-0.203
3206			Pb			18			0.354			0.267			-0.087			-0.245

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        480

---

3206			Zn			18			1.301			1.001			-0.3			-0.231
3207			Ag			10			23.048			14.677			-8.371			-0.363
3207			Au			10			0.948			0.886			-0.062			-0.065
3207			Cu			10			0.088			0.078			-0.011			-0.119
3207			Fe			10			25.692			24.156			-1.535			-0.06
3207			Pb			10			0.528			0.234			-0.294			-0.556
3207			Zn			10			1.044			0.485			-0.559			-0.536
3222			Ag			20			20.483			24.554			4.07			0.199
3222			Au			20			0.451			0.391			-0.061			-0.135
3222			Cu			20			0.064			0.067			0.003			0.053
3222			Fe			20			14.246			11.592			-2.654			-0.186
3222			Pb			20			0.734			0.772			0.039			0.053
3222			Zn			20			1.599			1.18			-0.419			-0.262
3223			Ag			16			16.129			20.761			4.632			0.287
3223			Au			16			0.596			0.767			0.172			0.288
3223			Cu			16			0.037			0.04			0.003			0.085
3223			Fe			16			12.905			13.332			0.427			0.033
3223			Pb			16			0.685			0.753			0.068			0.099
3223			Zn			16			1.398			1.172			-0.226			-0.161
3224			Ag			13			22.829			17.866			-4.963			-0.217
3224			Au			13			0.591			0.349			-0.242			-0.409
3224			Cu			13			0.062			0.082			0.019			0.31
3224			Fe			13			15.994			11.613			-4.381			-0.274
3224			Pb			13			0.571			0.386			-0.185			-0.324
3224			Zn			13			1.752			1.235			-0.517			-0.295
3225			Ag			3			20.595			20.495			-0.101			-0.005
3225			Au			3			0.942			0.859			-0.083			-0.088
3225			Cu			3			0.115			0.117			0.002			0.02
3225			Fe			3			11.259			11.94			0.681			0.06
3225			Pb			3			1.477			1.413			-0.063			-0.043
3225			Zn			3			0.891			0.969			0.078			0.088
3226			Ag			19			17.715			15.868			-1.848			-0.104
3226			Au			19			0.376			0.341			-0.035			-0.092
3226			Cu			19			0.083			0.056			-0.027			-0.328
3226			Fe			19			11.317			9.155			-2.161			-0.191
3226			Pb			19			0.741			0.531			-0.21			-0.283
3226			Zn			19			1.142			1.02			-0.122			-0.107
3227			Ag			8			32.543			30.644			-1.9			-0.058
3227			Au			8			1.068			0.789			-0.279			-0.261
3227			Cu			8			0.05			0.052			0.001			0.023
3227			Fe			8			13.363			12.367			-0.996			-0.075
3227			Pb			8			0.288			0.193			-0.095			-0.329
3227			Zn			8			0.948			0.589			-0.359			-0.379
3230			Ag			3			24.961			28.679			3.719			0.149
3230			Au			3			0.371			0.354			-0.017			-0.045
3230			Cu			3			0.026			0.02			-0.006			-0.223
3230			Fe			3			8.5			7.062			-1.437			-0.169
3230			Pb			3			0.638			0.585			-0.052			-0.082
3230			Zn			3			0.903			0.809			-0.094			-0.104
3291			Ag			92			44.288			35.786			-8.502			-0.192
3291			Au			92			0.931			1.01			0.079			0.085

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        481

---

3291			Cu			92			0.056			0.054			-0.002			-0.027
3291			Fe			92			18.019			17.327			-0.692			-0.038
3291			Pb			92			0.593			0.504			-0.089			-0.149
3291			Zn			92			1.124			1.014			-0.11			-0.098
3292			Ag			28			38.833			23.427			-15.406			-0.397
3292			Au			28			0.864			0.817			-0.048			-0.055
3292			Cu			28			0.046			0.037			-0.009			-0.193
3292			Fe			28			17.075			17.816			0.741			0.043
3292			Pb			28			0.431			0.262			-0.169			-0.392
3292			Zn			28			1.639			0.707			-0.931			-0.568
3293			Ag			41			24.417			20.73			-3.687			-0.151
3293			Au			41			0.541			0.534			-0.007			-0.013
3293			Cu			41			0.035			0.031			-0.005			-0.135
3293			Fe			41			16.022			13.445			-2.578			-0.161
3293			Pb			41			0.516			0.481			-0.035			-0.067
3293			Zn			41			1.051			0.8			-0.251			-0.239
3294			Ag			3			41.597			32.508			-9.089			-0.219
3294			Au			3			1.367			1.237			-0.13			-0.095
3294			Cu			3			0.091			0.054			-0.037			-0.406
3294			Fe			3			31.214			20.491			-10.722			-0.344
3294			Pb			3			0.056			0.047			-0.009			-0.162
3294			Zn			3			0.183			0.278			0.095			0.516
3295			Ag			33			34.646			37.046			2.4			0.069
3295			Au			33			0.328			0.292			-0.035			-0.108
3295			Cu			33			0.092			0.089			-0.003			-0.035
3295			Fe			33			16.446			16.928			0.481			0.029
3295			Pb			33			0.807			0.671			-0.137			-0.169
3295			Zn			33			1.368			1.399			0.031			0.022
3301			Ag			58			77.165			78.11			0.945			0.012
3301			Au			58			0.598			0.587			-0.011			-0.018
3301			Cu			58			0.098			0.081			-0.017			-0.172
3301			Fe			58			13.877			13.006			-0.871			-0.063
3301			Pb			58			0.559			0.471			-0.088			-0.158
3301			Zn			58			0.86			0.848			-0.012			-0.014
3402			Ag			5			11.989			10.624			-1.365			-0.114
3402			Au			5			0.329			0.386			0.057			0.172
3402			Cu			5			0.027			0.027			-0.001			-0.026
3402			Fe			5			11.929			11.712			-0.217			-0.018
3402			Pb			5			0.158			0.149			-0.01			-0.062
3402			Zn			5			0.474			0.529			0.055			0.116
3409			Ag			8			18.627			25.499			6.872			0.369
3409			Au			8			1.29			0.985			-0.305			-0.237
3409			Cu			8			0.018			0.023			0.005			0.296
3409			Fe			8			11.919			13.706			1.787			0.15
3409			Pb			8			0.304			0.233			-0.071			-0.234
3409			Zn			8			1.137			1.321			0.183			0.161
3410			Ag			6			28.02			12.419			-15.601			-0.557
3410			Au			6			0.659			0.839			0.18			0.274
3410			Cu			6			0.033			0.024			-0.008			-0.257
3410			Fe			6			19.76			17.426			-2.333			-0.118
3410			Pb			6			0.208			0.108			-0.1			-0.48

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        482

---

3410			Zn			6			1.482			0.852			-0.63			-0.425
3411			Ag			10			17.61			15.394			-2.216			-0.126
3411			Au			10			0.638			0.715			0.077			0.121
3411			Cu			10			0.026			0.024			-0.001			-0.049
3411			Fe			10			12.875			12.856			-0.019			-0.001
3411			Pb			10			0.799			0.491			-0.308			-0.386
3411			Zn			10			1.38			1.273			-0.106			-0.077
3508			Ag			7			20.82			25.177			4.358			0.209
3508			Au			7			0.883			1.049			0.167			0.189
3508			Cu			7			0.017			0.018			0.001			0.048
3508			Fe			7			17.48			18.304			0.824			0.047
3508			Pb			7			0.587			0.753			0.166			0.283
3508			Zn			7			1.578			1.656			0.078			0.049
3600			Ag			20			24.782			18.132			-6.649			-0.268
3600			Au			20			0.506			0.492			-0.014			-0.028
3600			Cu			20			0.024			0.03			0.007			0.285
3600			Fe			20			11.078			10.522			-0.556			-0.05
3600			Pb			20			0.789			0.536			-0.253			-0.321
3600			Zn			20			1.871			1.446			-0.425			-0.227
4001			Ag			3			483.308			535.819			52.511			0.109
4001			Au			3			0.071			0.043			-0.028			-0.392
4001			Cu			3			2.447			1.795			-0.652			-0.266
4001			Fe			3			5.155			3.459			-1.696			-0.329
4001			Pb			3			0.094			0.114			0.02			0.21
4001			Zn			3			0.088			0.097			0.009			0.101
4002			Ag			3			81.507			38.259			-43.248			-0.531
4002			Au			3			0.01			0.01			0			-0.002
4002			Cu			3			0.158			0.109			-0.049			-0.312
4002			Fe			3			3.666			2.996			-0.67			-0.183
4002			Pb			3			4.974			2.236			-2.738			-0.551
4002			Zn			3			6.715			4.543			-2.172			-0.323
4003			Ag			2			153.304			125.18			-28.124			-0.183
4003			Au			2			0.141			0.088			-0.053			-0.377
4003			Cu			2			0.594			0.391			-0.203			-0.342
4003			Fe			2			4.598			4.878			0.28			0.061
4003			Pb			2			7.782			5.972			-1.809			-0.233
4003			Zn			2			24.913			32.252			7.34			0.295

AYA GOLD & SILVER INC., BOUMADINE POLYMETALLIC PROJECT        483