Exhibit 96.7

 

  

Page i

 

SK-1300 Technical Report Summary, Initial Assessment on the Era Dorada Gold Project,

Jutiapa, Guatemala

 

 
GE21 Project No.: 250410
Effective date: December 31, 2024
Issue date: June 20, 2025
Version: Rev 01
Work directory: S:\Projetos\Aura\250410-InitAss-Cerro-Blanco
Copies: Aura Minerals Inc.
GE21 Consultoria Mineral Ltda.

 

 

 

 

 

 

 

 

Review Description Author(s) Date
01 Inclusion of cash flow table HLC 06/20/2025
       
       
       

 

Aura Minerals Inc. | Era Dorada Gold Project

SK-1300 Technical Report Summary – Initial Assessment

June, 2025
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DATe and signature

 

This report, entitled “SK-1300 Technical Report Summary, Initial Assessment on the Era Dorada Gold Project, Jutiapa, Guatemala”, having an effective date of December 31, 2024, was prepared by GE21 Consultoria Mineral Ltda. on behalf of Aura Minerals Inc., and signed.

 

Dated at Belo Horizonte, Brazil, on June 6, 2025.

 

 

 

 

 

 

/s/ Porfirio Cabaleiro Rodriguez

 

Porfirio Cabaleiro Rodriguez

 

 

/s/ Homero Delboni

 

Homero Delboni

 

 

/s/ Garth Kirkham

 

Garth Kirkham

 

 

 

 

 

 

 

 

 

Aura Minerals Inc. | Era Dorada Gold Project

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June, 2025
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NOTICE

 

 

GE21 Consultoria Mineral Ltda. prepared this Technical Report Summary, in accordance with S-K 1300 guidelines, for Aura Minerals Inc. The quality of information, conclusions and estimates contained herein is based on: (i) information available at the time of preparation; (ii) data supplied by outside sources, and (iii) the assumptions, conditions, and qualifications set forth in this report.

 

The Report is prepared in accordance with guidelines for mining property disclosure requirements provided in United States Securities and Exchange Commission (SEC) Regulation S-K, Subpart 1300 [S-K 1300]. Except for the purposes legislated under securities law, any other use of this report by any third party is at that party’s sole risk.

 

Before this report, this deposit was named “Cerro Blanco.” Aura Minerals Inc. has renamed it to “Era Dorada”. Therefore, all previous studies that were used as the basis for information in this report refer to the former name Cerro Blanco.

 

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UNITS, SYMBOLS, AND ABBREVIATIONS

 

Units and Symbols
% Percentage
Inche
°C Celsius
Au Gold
Au g/t grams of gold per tonne
CDN$ Canadian dollars
cm Centimetre
E East
g/m Gallon per Minute
g/t grams per tonne
Ga Gigaannum
Ha hectare(s)
hp horse power
hr hour
k thousand
k$ thousands of dollar
kg kilogram
km kilometre
kt thousands of tonnes
kV kilovolt
l litre
m metre
M million
m³/h cubic metre per hour
mg milligram
Mt million tonnes
Mtpa million tonnes per annum
NE northeast
NW northwest
Oz ounce
t/h tonnes per hour
tpd tonnes per day
USD United States dollar ($)
V volts
w/v weight by volume

  

 

Abbreviations
3D Three Dimensional
AA Atomic Absorption
AARL Anglo American Research Laboratories
AHD Average Hauling Distance
AI Abrasion Index
AMPRD Absolute Mean Paired Relative Difference
ANM National Mining Agency of Brazil
ASL Above Sea Level
BWI - Bond Work Index
CA Certificate of Authorization
CFEM Financial Compensation for Exploitation of Mineral Resources
CFR Code of Federal Regulations
CoG Cut-off Grade
CRM Certified Reference Material
Cum Cumulative
DDH Diamond Drill Hole

 

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Abbreviations
DGPS Differential Global Positioning System
DWT Drop Weight Test
EIA Environmental Impact Assessment
Esp Sphalerite
FA Fire Assay
FS Feasibility Study
GE21 GE21 Consultoria Mineral Ltda.
GPS Global Positioning System
GRG Gravity Recoverable Gold Tests
IBGE Brazilian Institute of Geography and Statistics
ICU Intensive Cyanidation Unity
IRR Internal Rate of Return
JV Joint Venture
LOM Life of Mine
LP Preliminary License
NPV Net Present Value
NSR Net Smelter Revenue
P80 Passing 80%
QA/QC Quality Assurance and Quality Control
QP Qualified Person
ROM Run of Mine
SEC Securities and Exchange Commission
SPI SAG power index and
SR Stripping Ratio

 

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TABLE OF CONTENTS

 

1 EXECUTIVE SUMMARY 1
  1.1 Introduction 1
  1.2 Reliance and Other Experts 1
  1.3 Property Description and Location 2
  1.4 Accessibility, Climate, Local Resources, Infrastructure, Physiography and Socio-Economic Context 2
  1.5 History 3
  1.6 Exploration 4
  1.7 Sample Preparation & Data Verification 5
  1.8 Mineral Processing and Metallurgical Testing 5
  1.9 Mineral Resource Estimate 6
  1.9.1 Methodology 6
  1.10 Mineral Reserve Estimate 7
  1.13 Mining Methods 8
  1.14 Process and Recovery Methods 8
  1.15 Infrastructure 9
  1.16 Market Studies 12
  1.17 Environmental Studies, Permitting, and Plans, Negotiations, or Agreements with Local Individuals or Groups 12
  1.17.1 Introduction 12
  1.17.2 Environmental Management and Permitting 12
  1.17.3 Water Management 12
  1.17.4 Waste Rock and Tailings Management 13
  1.17.5 Flora and Fauna 13
  1.17.6 Cultural and Archeological Resources 13
  1.17.7 Environmental Monitoring 13
  1.17.8 Environmental Management Plan 13
  1.17.9 Social Management 13
  1.17.10 Mine Closure 14

  

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  1.18 Capital and Operating Costs 14
  1.19 Economic Analysis 15
  1.20 Conclusions 16
  1.20.1 Risks 16
  1.21 Recommendations 17
2 INTRODUCTION 19
  2.1 Qualified Persons 20
  2.2 Site Visits and Scope of Personal Inspection 21
  2.3 Effective Date and Sources of Information 21
3 PROPERTY DESCRIPTION 23
  3.1 Property Location 23
  3.2 Property Description and Tenure 24
  3.3 Royalties 26
  3.4 Environmental 26
4 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE, AND PHYSIOGRAPHY 27
  4.1 Access 27
  4.2 Climate 27
  4.3 Physiography 28
  4.4 Local Resources & Infrastructure 29
5 HISTORY 31
  5.1 Data Validation History 32
  5.2 Historic Resources 34
  5.3 Historic Reserves 36
6 GEOLOGICAL SETTING, MINERALIZATION, AND DEPOSIT 38
  6.1 Introduction 38
  6.2 Regional Geology of Southern Guatemala 38
  6.3 Local Geology 41
  6.3.1 Lithology 42
  6.3.2 Structure 48

 

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  6.4 Deposit Geology 56
  6.5 Deposit Type 56
  6.6 Era Dorada Deposit Geology 59
  6.7 Mineralization 59
  6.7.1 Vein Zones 60
  6.7.2 Disseminated Mineralization 64
  6.7.3 Hydrothermal Alteration 65
7 EXPLORATION 68
  7.1 Goldcorp & Glamis Drilling (Pre-2017) 70
  7.2 Data Validation 71
  7.3 Bluestone Drilling (2017-2021) 73
  7.4 Significant Assay Results 74
8 SAMPLE PREPARATION, ANALYSES AND SECURITY 78
  8.1 Sampling Method & Approach 78
  8.1.1 Sampling Preparation, Analyses & Security (prior to November 2006) 78
  8.1.2 Sample Preparation, Analyses & Security (Goldcorp 2010 through 2012) 79
  8.1.3 Sampling Preparation, Analyses & Security (Bluestone 2017 to 2021) 80
  8.2 Quality Assurance & Quality Control 83
  8.2.1 QA/QC Performance & Discussion for Samples prior to 2017 83
  8.2.2 QA/QC Performance & Discussion of Results (Bluestone 2017 to 2021) 84
9 DATA VERIFICATION 88
  9.1 Geology, Drilling & Assaying 88
10 MINERAL PROCESSING AND METALLURGICAL TESTING 90
  10.1 Introduction 90
  10.2 Selected Testworks 91
  10.2.1 KCA (2012) – Leach Tests 91
  10.2.2 Phillips Enterprises (2011) – Comminution Tests 91
  10.2.3 Pocock Industrial (2011) – Solid / Liquid Separation Tests 92
  10.2.4 BaseMet (2018) – Chemical Assays 92
  10.2.5 BaseMet (2018) – Gravity Concentration 93

 

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  10.2.6 BaseMet (2018) – Leach Tests 93
  10.2.7 BaseMet (2018) – Cyanide Destruction Tests 96
  10.3 Summary and Conclusions 97
11 MINERAL RESOURCE ESTIMATES 99
  11.1 Introduction 99
  11.2 Data 100
  11.3 Data Analysis 101
  11.4 Geology & Domain Model 105
  11.5 Composites 108
  11.5.1 High-Grade Composite Analysis 111
  11.5.2 Low-Grade Composite Analysis 114
  11.6 Evaluation of Outlier Assay Values 117
  11.7 Specific Gravity Estimation 120
  11.8 Variography 121
  11.9 Block Model Definition 124
  11.10 Resource Estimation Methodology 125
  11.11 Mineral Resource Classification 126
  11.12 Stockpile Resources 128
  11.13 Mineral Resource Estimate 130
  11.14 Sensitivity of the Block Model to Selection Cut-off Grade 136
  11.15 Resource Validation 137
  11.16 Discussion with Respect to Potential Material Risks to the Resources 137
12 MINERAL RESERVE ESTIMATES 138
13 MINING METHODS 139
  13.1 Introduction 139
  13.2 Deposit Characteristics 139
  13.3 Geotechnical Analysis and Recommendations 140
  13.3.1 Rock Mass Characterization 140
  13.3.2 Geotechnical Domains and Rock Mass Properties 140
  13.3.3 In-situ Stresses 143

 

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  13.3.4 Empirical Stope Design Analysis 144
  13.3.5 Estimates of Unplanned Dilution 144
  13.3.6 Backfill Strength Requirements 145
  13.3.7 Ground Support 146
  13.4 Hydrogeology Analysis and Recommendations 147
  13.4.1 Evaluation of Dewatering Rates and Number of Locations 148
  13.4.2 Injection Wells 154
  13.5 Mining Methods 156
  13.5.1 Longhole Mining 156
  13.5.2 Mechanized Cut-and-Fill 158
  13.6 Mine Design 159
  13.6.1 Design and Optimization 159
  13.6.2 Access 161
  13.6.3 Development Types 161
  13.7 Mine Services 164
  13.7.1 Mine Ventilation 164
  13.7.2 Water Supply 165
  13.7.3 Dewatering 165
  13.8 Unit Operations 166
  13.8.1 Drilling 166
  13.8.2 Blasting 166
  13.8.3 Ground Support 166
  13.8.4 Mucking 167
  13.8.5 Hauling 167
  13.8.6 Backfill 167
  13.9 Mine Equipment 167
  13.10 Mine Personnel 168
  13.11 Mine Production Schedule 169
14 PROCESSING AND RECOVERY METHODS 172
  14.1 Description of the Process Plant 172

 

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  14.2 Design Criteria 173
  14.3 Process Plant Description 174
  14.3.1 Crushing 174
  14.3.2 Grinding 174
  14.3.3 Gravity Concentration and Intensive Leaching 174
  14.3.4 Pre-Leach Thickening 175
  14.3.5 Pre-Oxidation 175
  14.3.6 Leaching 175
  14.3.7 Carbon in Pulp – CIP 175
  14.3.8 Carbon Acid Wash, Elution and Regeneration (Carbon Processing) 176
  14.3.9 Electrowinning and Refining 177
  14.3.10 Cyanide Destruction 177
  14.3.11 Tailing Thickening and Filtering Circuit 177
  14.4 Reagent Handling, Storage and Preparation System 177
  14.5 Utilities and Water 178
  14.5.1 Air Supply / Oxygen 178
  14.5.2 Water Supply 179
15 INFRASTRUCTURE 180
  15.1 General 180
  15.2 General Site Layout 180
  15.3 Site Access Road 182
  15.4 Security 182
  15.5 Power Supply and Distribution 183
  15.5.1 Emergency Power 183
  15.5.2 Construction Power 183
  15.6 Process Plant 183
  15.7 Dewatering and Reinjection 184
  15.8 Truck shop, Warehouse, Mine Dry and Administration Buildings 184
  15.9 On-Site Water Tanks 184
  15.10 Bulk Fuel Storage and Delivery 185

 

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  15.11 Haul Roads 185
  15.12 Communications / IT 185
  15.13 First Aid / Emergency Services 185
  15.14 Explosives Storage and Magazines 186
  15.15 Sewage Treatment 186
  15.16 Surface Water Management 186
  15.17 Fresh Water Supply 187
  15.18 Water Treatment Infrastructure 187
  15.19 Tailings Management Facility 188
  15.19.1 Drystack Tailings Facility 188
  15.19.2 Subsurface Preparation 189
  15.19.3 Seepage Collection System 189
  15.20 Waste Rock Facility 190
16 MARKET STUDIES 192
  16.1 Gold Market 192
  16.1.1 Gold Price 192
  16.2 Silver Market 192
  16.2.1 Silver Price 193
17 ENVIRONMENTAL STUDIES, PERMITTING, AND PLANS, NEGOTIATIONS, OR AGREEMENTS WITH LOCAL INDIVIDUALS OR GROUPS 194
  17.1 Introduction 194
  17.2 Environmental Impact Assessment and Permitting 194
  17.2.1 EIA Areas of Influence 194
  17.2.2 Permitting 195
  17.3 Water Resources 196
  17.3.1 Water Quality 196
  17.3.2 Water Management 197
  17.4 Waste Rock and Tailings Management 197
  17.4.1 Waste Rock 197
  17.4.2 Tailings 197

 

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  17.4.3 Geochemistry Testwork 198
  17.5 Solid Waste Management 199
  17.5.1 Non-Hazardous Solid Waste 199
  17.5.2 Solid Hazardous Waste 199
  17.6 Flora and Fauna 199
  17.7 Cultural and Archeological Resources 200
  17.8 Environmental Monitoring 200
  17.9 Environmental Management Plan 200
  17.10 Social Management 200
  17.11 Mine Closure 201
  17.11.1 Underground Mine 201
  17.11.2 Process Plant 201
  17.11.3 Administration Offices and Ancillary Buildings 202
  17.11.4 Dry Stack Tailings Facility (DSTF) 202
  17.11.5 Waste Rock Facility 202
  17.12 Potential Risks and Mitigation Actions 202
  17.12.1 Permitting 202
  17.12.2 Tailings and Waste Rock 203
  17.12.3 Socio-Political 203
18 CAPITAL AND OPERATING COSTS 204
  18.1 Capital Cost Estimate 204
  18.1.1 Capital Cost Summary 204
  18.1.2 Capital Cost Profile 205
  18.1.3 Key Estimation Assumptions 206
  18.1.4 Key Estimation Parameters 206
  18.1.5 Basis of Estimate 206
  18.1.6 Indirect Cost Estimate 209
  18.1.7 Owner Cost Estimate. 210
  18.1.8 Closure Cost Estimate 212
  18.1.9 Contingency 212

 

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  18.1.10 Capital Estimate Exclusions 212
  18.2 Operating Cost Estimate 213
  18.2.1 Operation Labour 214
  18.2.2 Mining Operating Cost Estimate 215
  18.2.3 Processing Operating Cost 215
  18.2.4 General and Administration Operating Cost Estimate 216
19 ECONOMIC ANALYSIS 218
  19.1 Methodology 219
  19.2 Gold and Silver Prices 219
  19.3 Mine Production 219
  19.4 Plant Production 219
  19.5 Revenue 220
  19.6 Total Operating Cost 220
  19.7 Royalty Rights 220
  19.8 Capital Expenditure 221
  19.8.1 Initial Capital 221
  19.8.2 Sustaining capital 221
  19.8.3 Remediation and Closure Capital 221
  19.9 Total All in Sustaining Cost 221
  19.10 Working Capital 222
  19.11 Depreciation 222
  19.12 Exchange Rate Forecast 222
  19.12.1 Income Tax 222
  19.13 Discounted Cash Flow 223
20 ADJACENT PROPERTIES 226
21 OTHER RELEVANT DATA AND INFORMATION 227
22 INTERPRETATION AND CONCLUSIONS 228
  22.1 Geology & Mineral Resources 228
  22.1.1 Risks 229
  22.2 Mineral Processing and Metallurgical Testing and Processing and Recovery Methods 230

 

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  22.3 Mining Methods, Infrastructure, Capital and Operating Costs 230
  22.3.1 Risks 230
  22.4 Environmental Studies, Permitting, and Plans, Negotiations, or Agreements with Local Individuals or Groups 231
23 RECOMMENDATIONS 232
  23.1 Exploration, Geology & Resources 232
  23.2 Mineral Processing and Metallurgical Testing and Processing and Recovery Methods 232
  23.3 Mining Methods, Infrastructure, Capital and Operating Costs 232
  23.4 Environmental Studies, Permitting, and Plans, Negotiations, or Agreements with Local Individuals or Groups 233
24 REFERENCES 235
25 RELIANCE ON INFORMATION PROVIDED BY THE REGISTRANT 238

  

lIST OF TABLES

 

Table 1-1: Drilling summary 4
Table 1-2: Resource estimate using 2.25 g Au/t Cut-off 6
Table 1-3: Stockpile Resource estimate (Measured Resource) 7
Table 1-4: Estimated Project capital costs 14
Table 1-5: Estimated operating costs of the Project 15
Table 1-6: Simplified Discounted Cash Flow Results 16
Table 2-1: List of QPs and related responsibilities 21
Table 3-1: Coordinates of exploitation license “Era Dorada” 24
Table 3-2: Royalty assumptions 26
Table 5-1: Verification samples 32
Table 5-2: Drill hole collar survey (NAD 27 Zone 16N) 33
Table 5-3: Drill holes selected for data verification 33
Table 5-4: Indicated and Inferred Mineral Resource Estimate (2008) 34
Table 5-5: In-situ Mineral Resources (2014) 35
Table 5-6: Mineral Resource statement (2021) 36

 

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Table 5-7: Mineral Reserve estimate (2019) 36
Table 5-8: Cerro Blanco Gold Project open pit Mineral Reserve estimate (2022) 37
Table 7-1: Drilling summary 68
Table 7-2: Verification samples 72
Table 7-3: Drill hole collar survey (NAD 27 Zone 16N) 72
Table 7-4: Drill holes selected for data verification 72
Table 7-5: Gold & silver samples from the drill hole database 74
Table 8-1: Quantity of control samples by type (Bluestone 2017 to 2021) 84
Table 8-2: Summary of standards (Bluestone 2017 to 2021) 84
Table 8-3: Bluestone QA/QC sample insertion rates 85
Table 10-1: Head assays 91
Table 10-2: Gold extraction summary 91
Table 10-3: Comminution test results 92
Table 10-4: Head assays 93
Table 10-5: Gravity concentration results 93
Table 10-6: Bottle roll leach results 94
Table 10-7: Bottle roll leach results (CIP) 95
Table 10-8: Cyanide destruction results 97
Table 10-9: Preliminary recovery estimative 98
Table 11-1: Lithology units & codes 100
Table 11-2: Statistics for weighted gold & silver assays 101
Table 11-3: Statistics for weighted gold & silver assays for quaternary and cross-cutting rock types 102
Table 11-4: Statistics for weighted gold & silver assays for the Salinas Group rocks 102
Table 11-5: Statistics for weighted gold & silver assays for the Mita Group rocks 103
Table 11-6: Statistics for weighted gold & silver assays 104
Table 11-7: Vein groupings for derived for statistical, geostatistical and estimation 112
Table 11-8: Au composite statistics weighted by length for veins 113
Table 11-9: Silver composite statistics weighted by length for veins 114
Table 11-10: Numeric codes for lithologies 115

 

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Table 11-11: Gold composite statistics weighted by length for low-grade domains 116
Table 11-12: Silver composite statistics weighted by length for low-grade domains 117
Table 11-13: Cut grades for Au & Ag within vein domains 119
Table 11-14: Cut grades for Au & Ag within low-grade domains 119
Table 11-15: Cut vs. uncut comparisons for gold and silver composites within the high-grade vein domain groupings 119
Table 11-16: Cut vs. Uncut Comparisons for Gold and Silver Composites within the Salinas and Mita Domains 120
Table 11-17: SG zone assignments 120
Table 11-18: Geostatistical model parameters for gold by lithology unit 124
Table 11-19: Geostatistical model parameters for silver by lithology unit 124
Table 11-20: Stockpile Resource estimate (Measured Resource) 129
Table 11-21: Parameters used for stope optimization and cut-off grade 130
Table 11-22: Resource estimate using 2.25 g Au/t cut-off 131
Table 11-23: Sensitivity analyses of tonnage along with Au & Ag grades at various Au cut-off grades 136
Table 13-1: Mean rock mass properties by domain for 2011/2012 geotechnical core logging data 142
Table 13-2: Design rock mass quality ranges by geotechnical domain 144
Table 13-3: Estimates of unplanned dilution for longhole and cut-and-fill stopes by domain 145
Table 13-4: Required UCS for various stope widths 145
Table 13-5: Ground support recommendations for ore development 146
Table 13-6: Ground support recommendations for permanent development 147
Table 13-7: Wells planned and executed 153
Table 13-8: Cut-off grade calculation inputs 159
Table 13-9: Stope optimization parameters 160
Table 13-10: Underground dewatering system 166
Table 13-11: Mobile equipment fleet 167
Table 13-12: Underground mine operations personnel 168
Table 13-13: Mine production schedule 170
Table 14-1: Key process design criteria 173

 

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Table 14-2: Reagent consumption 178
Table 17-1: Main permit amendments & new permit required 196
Table 17-2: Current permits 196
Table 18-1: Capital cost summary 204
Table 18-2: Mine capital cost 207
Table 18-3: Contractor capital development rates 208
Table 18-4: Surface construction basis of estimate 208
Table 18-5: Indirect cost basis of estimate 209
Table 18-6: Closure estimate summary 212
Table 18-7: Breakdown of Estimated Operating Costs 213
Table 18-8: Main OPEX Component Assumptions 214
Table 18-9: Main OPEX Component Assumptions 214
Table 18-10: Underground Mine Operating Costs 215
Table 18-11: Process Operating Costs 215
Table 18-12: General and administration (G&A) operating cost summary 216
Table 18-13: G&A Labour Requirements & Costs 216
Table 19-1: Summary of key financial results 218
Table 19-2: Revenue composition 220
Table 19-3: Detailed operating costs 220
Table 19-4: All in sustaining costs composition 221
Table 19-5: Working capital periods 222
Table 19-6: Simplified Discounted Cash Flow

223

Table 19-7: Simplified Discounted Cash Flow (continued)

224

Table 19-8: Simplified Discounted Cash Flow Results

225

Table 22-1: Resource Estimate using 2.25 g Au/t Cut-off 229
Table 22-2: Stockpile Resource Estimate (Measured Resource) 229

 

LIST OF FIGURES

 

Figure 3-1: Project location map 23
Figure 3-2: Location of Mineral Resources relative to property boundary 24
Figure 3-3: Era Dorada exploitation license coordinates 25
Figure 4-1: Typical landscape in the Project area, looking South 28
Figure 4-2: Population centers near the Project area 30

 

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Figure 5-1: Example of XY scatter plot for hole CB34 34
Figure 6-1: Location of Era Dorada and other deposits in the Central American Volcanic 38
Figure 6-2: Regional structural map of Guatemala 39
Figure 6-3: Geological map of Era Dorada 42
Figure 6-4: Lithostratigraphy & lithology at Era Dorada 43
Figure 6-5: Examples of andesitic lapilli tuff (Mcv) 44
Figure 6-6: Examples of limestones (Mls) 45
Figure 6-7: Silicified reed fragments 46
Figure 6-8: Example Drill log from the Salinas Group 47
Figure 6-9: Recent travertine exposure 48
Figure 6-10: Simplified west-west cross-section across Era Dorada 49
Figure 6-11: East-west cross-section of the South zone, Era Dorada looking North 50
Figure 6-12: Stereograms (equal area) showing poles & great circles for faults & veins 52
Figure 6-13: Photographs with sketches of veins exposed underground 53
Figure 6-14: Annotated, vertical east-west cross-section across the south ramp (looking North) 54
Figure 6-15: Horizontal Slices at different elevations through Era Dorada 54
Figure 6-16: Stereograms for more detailed sub-areas in underground mapping 55
Figure 6-17: Generalized deposit model schematic 57
Figure 6-18: High-grade drill hole intercept hole CB20-430 – 144 g/t Au, 282 g/t Ag (227.3 to 228.9 m) 61
Figure 6-19: View of veins VN-05, 06, 07 in the north ramp underground workings 62
Figure 6-20: Examples of vein textures from Era Dorada 63
Figure 6-21: Example of geopetal structure 64
Figure 6-22: Salinas Unit – examples of disseminated mineralization rock types, Salinas Unit 65
Figure 6-23: Vertical alteration profile through Era Dorada 66
Figure 6-24: Examples Of Sealed, Silicified Fault Zones 67
Figure 7-1: Plan view of drill hole locations 69
Figure 7-2: Section View A-Aʹ (Azimuth 110°) 70
Figure 7-3: Section view B-Bʹ (azimuth 110°) 70
Figure 7-4: Example of XY scatterplot for hole CB34 73

 

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Figure 8-1: Example of core box photography 81
Figure 8-2: Example of underground channel sample 82
Figure 8-3: Batch plot of standard CDN-GS-6E 85
Figure 8-4: Plot of pulp & coarse reject duplicates (Bluestone 2017-2021) 86
Figure 8-5: Pulp & field blanks (Bluestone 2017 to 2021) 86
Figure 10-1: Effect of grind size on average gold extraction 94
Figure 10-2: Gold recovery as a function of residence time (CIP) 96
Figure 11-1: Plan view of drill holes 101
Figure 11-2: Box plot gold assays for the Salinas Group rocks 103
Figure 11-3: Box plot gold assays for the Mita Group rocks 104
Figure 11-4: Section view schematic of lithology for the Era Dorada Deposit 105
Figure 11-5: Plan view of drill holes & vein solids 106
Figure 11-6: South area section A-A’ view of drill holes, vein solids with Salinas and Mita Units 107
Figure 11-7: North area B-B’ section view of vein solids with Salinas and Mita Units 107
Figure 11-8: Histogram of assay interval lengths in metres 108
Figure 11-9: Histogram of assay interval lengths within veins in metres 109
Figure 11-10: Scatterplot of assay interval lengths within veins in metres versus gold grade 109
Figure 11-11: Histogram of gold composite grades (g/t) 110
Figure 11-12: Histogram of gold composite grades (g/t) with vein zones 110
Figure 11-13: Histogram of Silver Composite Grades (g/t) 111
Figure 11-14: Histogram of silver composite grades (g/t) with vein zones 111
Figure 11-15: Box plot of gold composites for veins 113
Figure 11-16: Box plot of silver composites for veins 114
Figure 11-17: Box plot of gold composites for low-grade domains 116
Figure 11-18: Box plot of silver composites for low-grade domains 117
Figure 11-19: Au cumulative frequency plot 118
Figure 11-20: Ag cumulative frequency plot 118
Figure 11-21: Au corellogram models 121
Figure 11-22: Ag corellogram models 122

 

 

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Figure 11-23: Ag correlogram models 123
Figure 11-24: Block model origin & orientation 125
Figure 11-25: Block model extents & dimensions 125
Figure 11-26: Plan view of stockpile, sample locations & domain solids 129
Figure 11-27: Plan view of gold block model with reasonable prospects optimized mine shapes with existing underground ramps 131
Figure 11-28: Plan view of Au within veins along with existing ramp development 132
Figure 11-29: Section view of Au south zone veins 132
Figure 11-30: Section view of Au block model north zone veins 133
Figure 11-31: Section view of Ag block model south zone veins 133
Figure 11-32: Section view of Ag block model north zone veins 134
Figure 11-33: Section view of Au block model south 134
Figure 11-34: Section view of Au block model north 135
Figure 11-35: Section view of Ag BLOCK MODEL NORTH 135
Figure 11-36: Section view of Ag block model south 136
Figure 13-1: Cross-section of geotechnical domain boundaries (looking North) 142
Figure 13-2: JDS (2018) geotechnical mapping Q’ values vs. RMR76 values 143
Figure 13-3: Existing location of portals, dewatering wells, monitoring wells and new dewatering well locations 149
Figure 13-4: Simulation hydrograph – south area and central area predicted dewatering at 2,500 and 3,500 g/m 151
Figure 13-5: Preliminary location of injection wells 155
Figure 13-6: Perspective view of a typical mining level 156
Figure 13-7: Longhole open stoping 157
Figure 13-8: Mechanized cut-and-fill 158
Figure 13-9: Mine long section 159
Figure 13-10: Drift profiles 163
Figure 13-11: Mine design plan view 163
Figure 13-12: Mine design long section (looking Northwest) 164
Figure 14-1: Overall Process Flowsheet – Era Dorada Project 173
Figure 15-1: Overall Mine Site 181

 

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Figure 15-2: Plant Site General Arrangement 182
Figure 15-3: Storm Water Management Infrastructure Surrounding the DSTF 187
Figure 15-4: DSTF Seasonal Material Placement Plan 188
Figure 15-5: DSTF Geotechnical Site Investigation Plan 189
Figure 15-6: DSTF Underdrain Plan Showing Starter Dam 190
Figure 15-7: WRF General Configuration Plan 191
Figure 16-1: Gold price behavior since 2000 192
Figure 16-2: Silver price behavior since 2000 193
Figure 17-1: EIA Areas of Influence 195
Figure 18-1: Distribution of initial capital cost 205
Figure 18-2: Distribution of sustaining capital cost 205
Figure 18-3: Capital cost profile 206
Figure 18-4: Operating Cost Distribution 214

 

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1EXECUTIVE SUMMARY

 

1.1Introduction

 

In January 2025, Aura Minerals Inc (“Aura” or “the Company”) completed the acquisition of the Era Dorada Gold Project—formerly “Cerro Blanco Gold Project”—and Mita Geothermal Project, located in Jutiapa, Guatemala, near the town of Asunción Mita and the border with El Salvador. The Era Dorada Project (“Era Dorada” or “the Project”) is 100% beneficially owned by Aura. Aura is a public, TSX-listed company trading under the symbol “ORA”, with its head office located at 78 SW 7th St., Miami, FL 33130, USA.

 

Aura commissioned GE21 Consultoria Mineral Ltda. (GE21) to prepare a Technical Report Summary (TRS) for the Project. The Mita Geothermal Project is not considered in this report.

 

This TRS, titled “SK-1300 Technical Report Summary, Initial Assessment on the Project, Jutiapa, Guatemala”, presents information in compliance with United States Securities and Exchange Commission’s (SEC) Modernized Property Disclosure Requirements for Mining Registrants as described in Subpart 229.1300 of Regulation S-K, Disclosure by Registrants Engaged in Mining Operations (S-K 1300) and Item 601 (b)(96) Technical Report Summary.

 

The purpose of this study is to document a Mineral Resource Estimate, mine design, and preliminary economics of the Project.

 

The Qualified Persons (QPs) responsible for this independent TRS are Mr. Porfirio Cabaleiro Rodriguez, Mr. Homero Delboni Jr., and Mr. Garth Kirkham. Neither GE21 nor the authors of this Independent TRS have had any material interest invested in Aura or any of its related entities. Their relationship with Aura is strictly professional, consistent with that held between a client and an independent consultant. This TRS was prepared in exchange for payment based on fees that were stipulated in a commercial agreement. Payment of these fees is not dependent upon the results of this TRS.

 

The effective date as it relates to the Initial Assessment is December 31, 2024, with the issue date of this TRS being June 6, 2025.

 

1.2Reliance and Other Experts

 

The information presented regarding the tenure, status, and work permitted by permit type is based on information published by the Ministerio de Energía y Minas (MEM).

 

This TRS has been reviewed for factual errors by Aura and all QPs. Any changes made as a result of these reviews did not involve any alteration to the conclusions made. Hence, the statements and opinions expressed in this TRS are given in good faith and in the belief that such statements and opinions are not false and misleading at the date of this TRS.

 

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1.3Property Description and Location

 

The Project is located in southeast Guatemala, in the Department of Jutiapa, approximately 160 km by road from the capital, Guatemala City, and approximately 9 km west of the border with El Salvador. The nearest town to the Project is Asunción Mita, situated approximately 7 km west of the Project. The exploitation license covers 15.25 km2 and lies entirely in the municipality of Asunción Mita.

 

The approximate center of the Project area is located at UTM coordinates X: 212,250 m E, Y: 1,587,250 m N, referenced to NAD27, Zone 16N. These coordinates correspond to the central portion of the mineral concession and are used for spatial reference in this report.

 

1.4Accessibility, Climate, Local Resources, Infrastructure, Physiography and Socio-Economic Context

 

Current road access to the site is via the Pan-American Highway (Highway CA1) through the town of Asunción Mita. Existing infrastructure is in place to provide year-round access to the site. The topography is relatively flat, with rolling hills.

 

The climate and vegetation at the Project site are typical of a tropical dry forest environment. The elevation is between 450 and 560 masl. The wet season is typically from May to October. The average annual rainfall is 1,350 mm. Daily temperature highs reach 41°C, and lows reach 10°C. The average annual pan evaporation rate is 2,530 mm, with an annual average humidity of 62%.

 

The Project is situated in proximity to several communities, the largest of which is Asunción Mita, with a population of approximately 18,500 people. The recently constructed La Barranca power substation is located a few kilometers south of Mita. The substation can supply up to 20 MW of power.

 

There is no record of any previous mining activity in the area; however, with the closure of Goldcorp’s Marlin Mine in late 2017, it is anticipated that a significant contingent of Guatemalan-trained labor will be available for employment at Era Dorada. As such, the Project intends to hire the majority of operations staff locally. It has allowed the Owner's budget to cover the cost of training programs.

 

A portion of the mine workforce is expected to live at the mine site in a purpose-built permanent camp, while employees living in the surrounding communities will provide their own transportation to and from the mine site. For employees residing in the wider Jutiapa region and areas beyond Asuncion Mita, the Company will provide transportation to and from the mine site from designated locations. There are several population centers near the Project site.

 

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1.5History

 

The Era Dorada property (formerly “Cerro Blanco”) was identified by Mar-West by a sampling of densely silicified boulders. In October 1998, Mar-West’s holdings in Honduras and Guatemala were purchased by Glamis Gold Ltd. In November 2006, Goldcorp Inc. became the sole proprietor of the Project through the purchase of Glamis Gold. Goldcorp undertook a comprehensive exploration program from 2006 to 2012, which included additional surface exploration, over 3.4 km of underground development, and 43,016 m of surface and underground drilling. On January 4, 2017, Bluestone agreed with Goldcorp to acquire 100% of the Project. On October 29, 2024, Aura purchased Bluestone Resources thereby acquiring 100% of the property.

 

As of the end of 2021, Bluestone had drilled approximately 267 holes for a total of 45,725 m on the Cerro Blanco property since the acquisition from Goldcorp. Geology & Mineralization

 

The Project is a classic hot springs-related, low-sulfidation epithermal gold-silver deposit comprising both high-grade vein and low-grade disseminated mineralization. The Cerro Blanco district is part of an active volcanic arc characterized by Miocene-Pliocene-aged bimodal volcanism that extends through El Salvador, Honduras, and Nicaragua.

 

High-grade mineralization is hosted in the Mita unit as two upward-flaring vein swarms comprising over 60 veins (North and South Zones) that converge downwards and merge into basal feeder veins. Low-grade disseminated and veinlet mineralization within and as halos around the high-grade veins is well documented in drilling since the discovery of the deposit. Most of the veins are blind to the surface and concealed by the syn-mineral Salinas Unit, a sub-horizontal sequence of volcanogenic sediments and sinter horizons approximately 100 m thick that form the low-lying hill at the Project. The Salinas cap rocks are host to low-grade mineralization associated with silicified conglomerates and contemporaneous dacite/rhyolite flow domes or cryptodomes.

 

Both high and low-angle banded crustiform/colloform chalcedony veins, locally with calcite replacement textures, make up the deposit, with bonanza-grade gold grades largely confined to the chalcedony-quartz veins, especially where adularia bands are prominent. High-grade mineralization occurs over a vertical profile of 400 m (150 to 450 masl). At depth, calcite-dominated veins form the limit to mineralization; nonetheless, very locally, high gold values are present in calcite-dominated veins and silicified structures containing only minor quartz veinlets.

 

The Salinas Group includes thin hot spring deposits, including sinters, which are genetically linked to underlying swarms of epithermal, gold-silver bearing quartz veins. The west and east sides of the Era Dorada ridge consist of flat agricultural plains characterized by Quaternary basalts, interbedded with boulder beds and sands. These rocks also appear down-faulted to lower elevations, implying major post-mineral extensional movements on such faults.

 

The current gold resource occurs under a small hill within an area of 400 m by 920 m. Gold-bearing structures in the Era Dorada area extend 2 km to the northwest of the gold deposit and occur largely confined within the hydrothermal alteration zone. The extensive drilling undertaken to date of the high-grade vein swarms and their surrounding low-grade mineralized

 

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envelopes and overlying mineralized cap rocks show impressive intercepts, including 203.8 m grading 2.3 g Au/t and 8.1 g Ag/t (CB20-420) and 87.2 m grading 5.9 g Au/t and 32.5 g Ag/t (UGCB18-89).

 

Vein textures suggest that gold and silver were introduced as one major event of multi-stage finely banded veining (originally amorphous silica) with subordinate bands of platy calcite, which is mostly pseudomorphed to cryptocrystalline silica phases. Repetitive “crack and seal” pulses and associated boiling/flashing events very close to the paleosurface are proposed as the main mechanisms for precious metal deposition. Very high-grade core intersections with coarser and more abundant sulfides, electrum, and free gold appear to represent an earlier series of events. Deportment studies indicate that approximately 99% of the gold occurs in electrum as free or exposed grains, with lesser amounts as native gold and kustelite. The lack of post-mineral structural displacement of veins and distribution of high grades over a +400 m vertical profile attest to the pristine nature of the veins at Era Dorada. The lack of inter-stage hydrothermal brecciation and coarse-grained primary quartz textures suggest that the mineralizing event was fairly short-lived and occurred very close to the paleosurface.

 

1.6Exploration

 

The Era Dorada property was identified by Mar-West by a sampling of densely silicified boulders. In October 1998, Mar-West’s holdings in Honduras and Guatemala were purchased by Glamis Gold Ltd. In November 2006, Goldcorp Inc. became the sole proprietor of the Project through the purchase of Glamis Gold. Goldcorp undertook a comprehensive exploration program from 2006 to 2012, which included additional surface exploration, over 3.4 km of underground development, and 43,016 m of surface and underground drilling. On January 4, 2017, Bluestone agreed with Goldcorp to acquire 100% of the Project and on October 29, 2024, Aura acquired Bluestone Resources.

 

As of the end of 2021, Bluestone had drilled approximately 267 holes for a total of 45,725 m on the Era Dorada property since the acquisition from Goldcorp. Table 1-1 summarizes the historical drilling on the property.

 

Table 1-1: Drilling summary

 

Year Company Holes Drilled Meters
1998 Mar-West 9 1,340
1999 Glamis 48 7,074
2000 Glamis 18 3,525
2002 Glamis 23 6,525
2004 Glamis 42 9,370
2005 Glamis 120 29,065
2006 Glamis 67 15,129
2007 Goldcorp 47 12,373
2008 Goldcorp 2 586
2009 Goldcorp 1 140
2010 Goldcorp 10 2,277
2011 Goldcorp 28 5,898
2012 Goldcorp 96 21,370

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Year Company Holes Drilled Meters
2017 Bluestone 8 2,324
2018 Bluestone 74 13,993
2019 Bluestone 61 8,403
2020 Bluestone 74 15,172
2021 Bluestone 50 5,833
Total 778 160,397

 

Source: Bluestone, 2021.

 

1.7Sample Preparation & Data Verification

 

Garth Kirkham, P. Geo., has been involved with the property since its acquisition in early 2017, when he performed the initial due diligence and authored the updated resource estimate for Bluestone. Mr. Kirkham first visited the property on May 8, 2017, to validate all aspects of the Project. The site visit included an inspection of the property, offices, underground vein exposures, core storage facilities, water treatment plant, and stockpiles, and a tour of major centers and the surrounding villages most likely to be affected by any potential mining operation.

 

Since 2017, Mr. Kirkham has visited the property numerous times for extended periods to develop and implement data-gathering and sampling methods and procedures. He also worked with Bluestone geologists to develop drill programs and supervise interpretation and modeling efforts, in addition to creating and implementing QA/QC procedures. Continued data validation and verification processes have not identified any material issues with the Era Dorada sample and assay data.

 

During Q3 and Q4 2020, the Era Dorada drill and assay database was moved to the AcQuire - GMSuite platform hosted by CSA Global, providing an enhanced and more secure standard of data management.

 

It is the opinion of the QP, Garth Kirkham, P. Geo., that the sampling preparation, security, analytical procedures, and quality control protocols used at Era Dorada are consistent with generally accepted industry best practices and are, therefore, reliable for resource estimation.

 

1.8Mineral Processing and Metallurgical Testing

 

Metallurgical test work was conducted on samples from the Era Dorada deposit (formerly named “Cerro Blanco”) between April 1999 and January 2012 by Kappes, Cassiday & Associates (KCA) and in 2018 by Base Metallurgical Laboratories Ltd. (BaseMet) in Kamloops, BC.

 

The test work programs included comminution testing, determination of head assays, grinding size assessments, gravity concentration, leach testing, tailings testing, and cyanide destruction.

 

Data obtained from both test work campaigns were used to estimate gold and silver recoveries, as well as to define the processing flowsheet configuration and process design criteria.

 

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For the global composite sample, the average recoveries obtained were 96% Au and 85% Ag.

 

1.9Mineral Resource Estimate

 

Era Dorada is a classic hot springs-related, low-sulfidation epithermal gold-silver deposit comprising both high-grade vein and low-grade disseminated mineralization. Most of the high-grade mineralization is hosted in the Mita unit as two upward-flaring vein swarms (north and south zones) that converge downwards and merge into basal feeder veins where drilling has demonstrated widths of high-grade mineralization (e.g., 15.5 m 21.4 g Au/t and 52 g Ag/t). Bonanza gold grades are associated with ginguru banding and carbonate replacement textures. Sulfide contents are low, typically < 3% by volume. Low-grade disseminated and veinlet mineralization in wall rocks around the high-grade veins is well documented in drilling since the discovery of the deposit, with grades typically ranging from 0.3 to 3.0 g/t Au.

 

The Salinas unit overlies the Mita rocks, a sub-horizontal sequence of volcanogenic sediments and sinter horizons approximately 100 meters thick, which form the low-lying hill at the Project. Low-grade disseminated, and veinlet mineralization within and as halos around the high-grade vein swarms is well documented in drilling since the discovery of the deposit, with grades typically ranging from 0.3 to 1.5 g Au/t. The overlying Salinas cap rocks are also host to low-grade mineralization associated with silicified conglomerates and rhyolite intrusion breccias.

 

Mineral exploration activities conducted at Era Dorada have been performed in accordance with S-K 1300.

 

1.9.1Methodology

 

The mineral resource estimate reported herein was prepared by Mr. Garth Kirkham, P. Geo. The mineral resources have been estimated in conformity with generally accepted CIM “Estimation of Mineral Resource and Mineral Reserves Best Practices.” There are 130,307 gold assays, totaling 153,078 m, which average 0.68 g/t, and 130,238 silver assays, totaling 153,003 m, which average 3.75 g/t. Bulk densities were assigned to individual rock types and assigned on a block-by-block basis using measurement data by lithology and mineralized veins.

 

The estimate was completed using MineSightTM software with a 3D block model (5 m x 5 m x 5 m). Interpolation parameters have been derived based on geostatistical analyses conducted on 1.5-meter composited drill holes. Block grades have been estimated using ordinary kriging (OK) methodology, and the mineral resources have been classified based on proximity to sample data and the continuity of mineralization in accordance with S-K 1300 requirements.

 

Table 1-2: Resource estimate using 2.25 g Au/t Cut-off

 

Resource Category Tonnes (kt) Au Grade (g/t) Ag Grade (g/t) Contained Gold (koz) Contained Silver (koz)
Measured          
Indicated 6,349 9.31 31.54 1,901 6,439
Measured & Indicated 6,349 9.31 31.54 1,901 6,439

 

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Resource Category Tonnes (kt) Au Grade (g/t) Ag Grade (g/t) Contained Gold (koz) Contained Silver (koz)
Inferred 605 6.02 19.68 117 383
   
Notes:

The mineral resource statement is subject to the following:

1.Mineral Resources are reported in in accordance with S-K 1300.

2.Mineral resource estimates have been prepared by Garth Kirkham, P.Geo., a Qualified Person as defined by SK-1300.

3.The Mineral Resource estimate is reported on a 100% ownership basis.

4.Underground mineral resources are reported at a cut-off grade of 2.25 g Au/t. Cut-off grades are based on a assumed metal prices of US$ 2,500/oz gold and US$ 28/oz silver, and assumed metallurgical recovery, mining, processing, and G&A costs.

5.Mineral Resources are reported without applying mining dilution, mining losses, or process losses.

6.Resources are constrained within underground shapes based on reasonable prospects of economic extraction, in accordance with SK-1300. Reasonable prospects for economic extraction were met by applying mining shapes with a minimum mining width of 2.0 m, ensuring grade continuity above the cut-off value, and by excluding non-mineable material prior to reporting.

7.Metallurgical recoveries reported as the average over the life of mine and are assumed to be 96% Au and 85% Ag, respectively.

8.Bulk density is estimated by lithology and averages 2.47, 2.57 and 2.54 g/cm3 for the Salinas, Mita and mineralized vein domains, respectively.

9.Mineral resources are classified as Indicated, and Inferred based on geological confidence and continuity, spacing of drill holes, and data quality.

10.Effective date of the mineral resource estimate is December 31, 2024.

11.Tonnage, grade, and contained metal values have been rounded. Totals may not sum due to rounding.

12.Mineral resources are not mineral reserves and do not have demonstrated economic viability.

Source: Kirkham, 2025.

 

In addition, there has been mixed grade material mined during the creation of the extensive, existing ramp network which has been stockpiled adjacent to North Ramp entrance. Table 1-3 shows the volume and tonnage based on an unconsolidated specific gravity of 2.0 gm/cm3 along with gold and silver grades and metal content. These resources are classified as measured.

 

Table 1-3: Stockpile Resource estimate (Measured Resource)

 

Volume (BCM) Mine (t) Au (g/t) Ag (g/t) Au (oz) Ag (oz)
14,863 29,726 5.35 22.59 5,108 21,590

 

Source: Kirkham, 2019.

 

1.10Mineral Reserve Estimate

 

Mineral resources are not Mineral Reserves and have no demonstrated economic viability. This initial assessment does not support an estimate of Mineral Reserves since a pre-feasibility or Feasibility Study is required for reporting mineral reserve estimates. This report is based on potentially mineable material (mineable tonnes and do not Reserve Estimate).

 

Mineable tonnages were derived from the resource model described in the previous section. Measured, Indicated, and Inferred resources were used to establish mineable tonnes.

 

Inferred mineral resources are considered too speculative geologically to have economic considerations applied to them that would enable them to be categorized as Mineral Reserves. There is no certainty that all or any part of the Mineral Resources or mineable tonnes would be converted into Mineral Reserves.

 

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1.13Mining Methods

 

High-grade mineralization at the Era Dorada deposit is hosted within laterally stacked, sub-parallel narrow veins that generally strike northeast, with average azimuth ranging from 25° to 50°. Vein dip varies, including both tabular and near-vertical geometries.

 

The average dip of high-grade structures is approximately 50° to 55°. The average vein thickness ranges from 2 to 10 m, with an average spacing of 8 m between parallel structures. The potentially mineable resource ranges from 50 m at the lowest levels to 300 m near the surface. The mineralized system comprises more than 50 modeled veins with variable geometry along both strike and dip.

 

Era Dorada is proposed to be mined as an underground operation using a combination of longhole stoping (LH) and mechanized cut-and-fill (MCF) mining methods, with cemented paste and rock backfill. A target production rate of 1,500 tpd is envisioned over a mine life of 17 years, which will extract 8.9 Mt of ore. LH stoping will account for about 77% of total production, and the remaining 23% will come from MCF and development. The Era Dorada deposit will be accessed from surface via a series of ramps, and all ore and waste rock will be trucked out of the mine. In addition to the four existing ventilation raises, two new raises will be required to circulate the required amount of air through the underground workings.

 

Dewatering, ventilation, and cooling are crucial aspects of mine design at Era Dorada. A series of existing and new surface dewatering wells will lower the water levels in the immediate mine area. Any remaining water underground will be captured and pumped to the surface through collection at underground sumps. For ventilation, the quantity of air required has been designed to dilute diesel particulate matter, reduce the air temperature from exposed rock, and maintain worker comfort. Mine air refrigeration will be used to maintain air temperatures in working areas below 28°C wet bulb.

 

Indicated and Inferred Mineral Resources were included in the mine design and schedule optimization process. The Indicated material accounts for 78% of LoM, while Inferred material accounts for 22%.

 

The mine production schedule is shown in Table 13-13.

 

1.14Process and Recovery Methods

 

The processing plant will process 1,500 tpd, consisting of the following unit operations:

 

·Crushing circuit.

 

·Grinding circuit to a nominal P80 of 0.053 mm.

 

·Gravity concentration and intensive leaching (ILR).

 

·Pre-leach thickening to 50% solids (w/w).

 

·2-hour pre-oxidation, 36-hour leaching, and 6-hours Carbon-in-Pulp (CIP).

 

·Carbon acid wash, elution, and regeneration.

 

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·Electrowinning and refining.

 

·Cyanide destruction.

 

·Tailings thickening, filtration, and disposal in the DSTF or underground as paste backfill.

 

The leach circuit will have a residence time of 36 hours. The sodium cyanide (NaCN) consumption is predicted to be in the range of 0.3 to 0.5 kg/t to maintain a cyanide concentration of 500 ppm. Cyanide will be destroyed using the SO2/Air process (Detox circuit). Tailings resulting from the Detox circuit will be transferred to a thickener, whose underflow will be pumped to the filtration circuit, where a horizontal vacuum filter will reduce the cake moisture to 18.6% (dry basis).

 

Most of the water consumed in the processing plant is designed to derive from recirculation within the industrial installation.

 

The main reagents to be used in the Era Dorada industrial plant are sodium cyanide, hydrated lime, lead nitrate sodium hydroxide, sodium metabisulfite, hydrochloric acid, copper sulfate pentahydrate, and flocculant.

 

1.15Infrastructure

 

The Project plans the installation of the following elements to support the mine and process facilities:

 

·5,5 km new site access road, including a 80 m long bridge;

 

·8.2 km new 69 kV power line;

 

·on-site substation (69 kV to 13.8 kV);

 

·water management facilities, including a flood protection levee, diversion channel, ditches, and collection ponds;

 

·process plant site pad and associated buildings;

 

·primary crusher pad;

 

·emergency power genset;

 

·communications system upgrade;

 

·rehabilitation of five existing dewatering wells;

 

·construction of eight new dewatering wells;

 

·construction of nine new reinjection wells;

 

·reagent warehouse and storage facilities;

 

·truck shop (existing facility to be used in pre-production, new shop to be constructed in Operating Year 1);

 

·fresh / fire water tank;

 

·process water tank;

 

·upgrade fuel storage facility;

 

·new helipad;

 

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·upgrade septic system for the upgrade for sewage management;

 

·solid waste disposal facility;

 

·dry stack tailings facility (DSTF);

 

·temporary waste rock storage facility;

 

·1.0 km North and South portal connector haul road;

 

·on-site access roads for plant and facilities;

 

·additional security facilities, including a site access control station.

 

The proposed site layout has been designed to support mining and plant operations while minimizing environmental and community impacts, reducing construction costs, ensuring secure access, and optimizing operational efficiency.

 

Existing Facilities: Administrative, technical, geology, environmental, and assay lab facilities are established and will remain operational. Logistics, security, equipment servicing, and sample processing infrastructure are also in place.

 

Access and Security: Current access via a gravel road with a 27-t bridge is insufficient. A new 5.5 km access road connecting directly to the Pan-American Highway with an 80 m bridge over El Achotal River will be built. The site entrance will feature a secure gate and access control. Fencing and security personnel will oversee site safety, with heightened measures in place during construction.

 

Power Supply: Power will be supplied via an 8.2 km, 69 kV transmission line from Energuate Barranca Honda Substation, stepping down to 13.8 kV and lower voltages as needed. Emergency power will utilize relocated diesel generators to support critical loads (~7 MVA). Temporary generators will serve construction needs.

 

Process Plant: Approximately 150 m x 70 m, including grinding, leaching, Merrill-Crowe, filtration, detoxification, reagent prep, dry stack tailings filtration, and electrical rooms. Enclosed milling and refinery facilities will be built to code; MCCs and control rooms will be preassembled where possible.

 

Water Management: Existing and new dewatering wells (totaling 24 wells) will manage groundwater inflows, with peak surface dewatering at ~795 m³/h, supplemented by underground sumps. A new cooling pond and expanded water treatment plant (capacity 341 m³/h) will treat mine water for process use and personnel facilities—potable water supplied by local bottled water vendors. Water reuse is emphasized through the use of reinjection wells and a zero-discharge strategy.

 

Maintenance Facilities: Mine truck shop near administration and North Portal with three service bays, welding/general shop, oil change bay, outdoor wash bay, parts warehouse, and offices—steel structure with 10-t crane.

 

Water Storage: New dual-purpose fresh/fire water tank (640,000 l) with minimum 470,000 l fire reserve; adjacent 170,000 l process water tank.

 

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Fuel Storage: Existing diesel tanks (2 × 37,500 l) expanded by one additional tank and containment area. Capacity supports 14 days of mobile equipment or 2 days of critical loads.

 

Access Roads: North and South portal roads widened to 22 m, with the north road extended to the dry stack tailings facility. Temporary construction roads developed as needed.

 

Communications: Existing tower supplemented with fiber optic cable alongside 69 kV power line. Dedicated underground mine communication system and handheld radios for mobile equipment and security.

 

Emergency Services: First aid clinic under construction with training room and medical storage. Ambulance and fire truck stationed near the process plant. Site-wide safety detectors and fire extinguishers installed.

 

Explosives Storage: Existing explosives magazine continues in use with a capacity for 75,000 kg explosives and 10,000 detonators; monthly deliveries planned; compliant safety berms.

 

Sewage Treatment: Septic systems with bio-reactors for sanitary waste; existing and new units for expanded facilities; treated wastewater separated before discharge.

 

Surface Water Management: Separation of contact and non-contact water; contact water reused or treated; non-contact runoff directed to sediment control ponds and discharge points. Stormwater managed by lined channels, ponds, culverts, and erosion controls designed for 100-year storm events.

 

Dry Stack Tailings Facility (DSTF): Designed for 3 million tonnes (1.9 million m³) of filtered tailings with centerline-raised embankment. Seasonal deposition for stability and runoff control. Tailings transported by haul trucks. Initial infrastructure includes impoundments, underdrains, geotextile liners, reclaim and stormwater ponds, and water recirculation systems.

 

Geotechnical Investigations: Extensive test pits, boreholes, SPT, and permeability testing confirmed subsurface conditions (colluvial/alluvial soils over residual sedimentary and volcanic materials) and supported foundation design and seismic performance.

 

Seepage System: Foundation underdrains and vertical decant towers collect seepage and runoff, directing water to reclaim and stormwater ponds with pumps returning water to the process plant, supporting zero-discharge management.

 

Waste Rock Facility (WRF): Located near the south portal, designed for 120,000 m³ temporary storage, primarily for underground backfill. Preliminary geochemical tests indicate low acid generation risk; further testing and detailed geotechnical studies planned.

 

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1.16Market Studies

 

Mineral Resources were estimated using a gold price of US$ 2,000 per troy ounce. Project economics were evaluated at a gold price of US$ 2,389 per troy ounce and a silver price of US$ 28.44 per troy ounce. All price assumptions are based on the long-term consensus forecast from over 20 investment banks.

 

1.17Environmental Studies, Permitting, and Plans, Negotiations, or Agreements with Local Individuals or Groups

 

1.17.1Introduction

 

Aura’s review shows that the Project has all necessary permits to proceed with the development of the underground mine and construction of the process facilities, subject to the future operation to adhere to the conditions of the existing permits.

 

1.17.2Environmental Management and Permitting

 

Environmental studies have been conducted at Era Dorada since the Project's inception. The Environmental Impact Assessment (EIA) was submitted and approved for an underground mine in 2007 by Guatemala’s Ministry of Environment and Natural Resources (MARN). However, the Project design changed since 2007 and requires permit amendments. Additionally, new baseline studies are necessary for infrastructure components such as power lines. The approved EIA includes an Environmental Management Plan (EMP), a Social Management Plan (SMP), and a Conceptual Mine Closure Plan, which have been reviewed and updated following international best practices.

 

1.17.3Water Management

 

The Project's water management infrastructure consists of a Water Treatment Plant (WTP), pipelines, settling ponds, wells, and cooling channels. Monitoring of surface and groundwater is conducted regularly, with compliance reports submitted to MARN. Naturally occurring metals such as aluminum and arsenic are found in local waters. The WTP, designed for arsenic removal, uses ferric salt co-precipitation and ensures compliance with Environmental Protection Agency (EPA) and Guatemalan standards.

 

·Surface Water Management: Runoff is classified as “contact” or “non-contact” water. Contact water undergoes treatment before being reused or discharged, while non-contact water is diverted and monitored.

 

·Groundwater Management: Dewatering of the mine is achieved through surface wells and underground sumps. Treated water is either reused or discharged into Quebrada Tempisque. Reinjection wells are used to manage groundwater.

 

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1.17.4Waste Rock and Tailings Management

 

·Waste Rock: Temporary storage of waste rock occurs for a maximum of one year before being returned to underground as backfill. Based on the limited exposure time and historical geochemistry testing, it was assumed that any potential acid generation would not have sufficient time to occur. Comprehensive Environmental monitoring will indicate and anticipate any potential acid generation to prevent impacts.

 

·Tailings: Tailings are dewatered through filtration before placement in the Dry Stack Tailings Facility (DSTF). The facility, designed to prevent environmental contamination, collects runoff water for treatment. Based on tests of geochemical characterization and consistent with the approved EIA, tailings are considered to be non-acid generating (NAG).

 

1.17.5Flora and Fauna

 

Baseline studies have been conducted in the region since 2007, documenting its biodiversity. Ongoing monitoring indicates minimal impact from the Project. The local ecosystem comprises subtropical and tropical dry forests, which support a diverse array of plant species. Wildlife monitoring shows stable populations of birds, reptiles, and aquatic fauna. Preventive conservation measures, including habitat relocation for threatened species, have been implemented.

 

1.17.6Cultural and Archeological Resources

 

A dedicated on-site team monitors the potential impact on cultural and archaeological artifacts. Pre-construction inspections and external expert consultations ensure compliance with relevant regulations. To date, no significant historical artifacts have been identified within the Project’s direct area of influence.

 

1.17.7Environmental Monitoring

 

The Project maintains 26 monitoring stations for water quality, six for air quality, and additional noise monitoring locations. Monthly reports are submitted to regulatory authorities. While air quality and noise monitoring were not included in the original 2007 baseline study, comparisons to EPA standards ensure compliance.

 

1.17.8Environmental Management Plan

 

The Environmental Management Plan (EMP) has been updated to incorporate lessons learned from a decade of on-site environmental data collection. The plan aligns with regulatory requirements and international best practices. It integrates corporate health, safety, and environmental programs, including emergency response strategies.

 

1.17.9Social Management

 

Aura prioritizes strong community relationships. The Project retains a comprehensive database of community engagement activities and sustainability initiatives. The updated Social

 

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Management Plan (SMP) incorporates IFC performance standards and includes mechanisms for communication, grievance handling, and community engagement. A Social Monitoring Committee (SMC) is being established to ensure transparency.

 

1.17.10Mine Closure

 

The approved EIA includes a conceptual mine closure plan, which was further refined.

 

Mine closure requirements include:

 

·Progressive underground backfilling of waste rock and tailings.

 

·Decommissioning of infrastructure while maintaining environmental safeguards.

 

·Long-term monitoring of water quality and ecosystem restoration.

 

·Revegetation using native plant species.

 

The Dry Stacking Tailings Facility (DSTF) will be constructed continuously over the life of the mine using the downstream construction method, so concurrent reclamation will not be possible. At the end of operations, exposed portions of the decant piping will be dismantled, and the decant pipes will be plugged below the final surface.

 

The surface of the DSTF will be contoured so that it will shed precipitation rather than impound it. Topsoil that is stockpiled from the DSTF footprint during construction will be spread over the surface of the DSTF. Native grass seed mixture will be planted to reduce erosion.

 

Total closure costs are estimated at $17.19 M and do not include contingencies.

 

1.18Capital and Operating Costs

 

LoM Project capital costs total $ 417 M, consisting of the following distinct phases:

 

·Pre-production capital costs – includes all costs to develop the property to a 1,500 tpd production. Initial capital costs total $ 263.6 M and are expended over a 23-month pre-production period on engineering, construction, and commissioning activities, as shown in Table 18.1: Estimated Pre-production Capital Costs.

 

·Sustaining capital costs – includes all costs related to the acquisition, replacement, or major overhaul of assets during the mine life required to sustain operations. Sustaining capital costs total $136.2 M and are expended in operating Year 1 through Year 24.

 

·Closure Costs – includes all costs related to the closure, reclamation, and ongoing monitoring of the mine, post operations. Closure costs total $ 17.2 M and are primarily incurred in Year 14, with costs extending into Year 17 for ongoing monitoring activities.

 

Table 1-4: Estimated Project capital costs

 

WBS DESCRIPTION Pre-Production
Cost USD ($M)
Sustaining
Cost/Closure USD
($M)
Project Total
Cost USD ($M)
Infrastructure 8.2 8.4 16.6
Power and Electrical 16.7 - 16.7
Water Management 16.1 24.7 40.8

 

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WBS DESCRIPTION Pre-Production
Cost USD ($M)
Sustaining
Cost/Closure USD
($M)
Project Total
Cost USD ($M)
Surface Operations 14.7 1.7 16.4
Mining 63.4 80.2 143.6
Process Plant 49.7 16.7 66.4
Construction Indirect 38.0 4.6 42.6
General Services – Owner’s Costs 21.3 - 21.3
Logistics/ Taxes/ Insurance 9.0 - 9.0
Pre- Production, Start-up & Comissioning 5.0 - 5.0
Contingency 21.9 - 21.9
Closure Costs - 17.2 17.2
TOTAL 263.6 153.5 417.0

 

Source: Aura, 2025.

 

The operating cost estimate in this study includes the costs to mine and process the mineralized material to produce doré, as well as site services to maintain the site and general and administrative expenses (G&A). These items total the Project’s operating costs and are summarized in Table 1-5. The target accuracy of the operating cost is -30% to +50 %. The operating cost estimate is broken into four major sections:

 

·Underground mining

 

·Processing

 

·Site services

 

·General and Administrative (G&A).

 

The total operating unit cost is estimated to be US$ 170/t processed. Average annual, total LOM, and unit operating cost estimates are summarized in Table 1-5. The unit rates in this table include tonnes mined during the pre-production period.

 

Table 1-5: Estimated operating costs of the Project

 

Operating Costs Avg Annual (M$) $/t processed LOM (M$)
Mining 38.70 100 890.01
Processing 12.38 32 284.80
Site Services 6.97 18 160.20
G&A 7.74 20 178.00
Total 65.78 170 1,513.01

 

Source: GE21, 2025.

 

1.19Economic Analysis

 

The economic analysis for the Era Dorada Project is based on Mineral Resource estimates, including the annual mine production schedule. As required under SK-1300, the results of this analysis should not be interpreted as demonstrating the economic viability of the project.

 

The outcome of the economic analysis is subject to known and unknown risks, uncertainties, and other factors that may cause actual results to differ materially from them. The information on which this analysis is based is listed below:

 

·Mineral Resource Estimates

 

·Assumed fixed exchange rate

 

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·Assumed known royalties

 

·Proposed mine production plan

 

·Projected mining and processing recovery rates

 

·Fixed installed processing plant capacity

 

·Assumptions on closure costs

 

·Assumptions on environmental, licensing, and social risks

 

·Changes in production costs relative to the assumptions

 

This analysis does not rely on:

 

·Unrecognized environmental risks

 

·Unanticipated recovery expenses

 

·Different geotechnical and/or hydrogeological considerations during mining

 

·Unexpected variations in the quantity of mineralized material, grade, metallurgical recovery efficiency, and plant recovery efficiency

 

·Accidents, labour disputes, and other mining industry risks

 

·Changes in tax rates

 

·Assumptions of commercial discounts that are not foreseen in the financial analysis

 

The economic results are presented in Table 1-6.

 

Table 1-6 - Simplified Discounted Cash Flow Results

 

Discount Rate (%) .5%
NPV – After Tax (M US$) 485.49
IRR – After Tax (%) 23.8%
Payback - After Tax (years) 3.75

 

Source: GE21, 2025.

 

1.20Conclusions

 

The Project outlines a conceptual mine plan involving the extraction of 8.9 Mt of ROM over a 17-year LoM, with a production rate of 1,500 tpd. The selected underground mining method is suitable for ensuring a stable and consistent mill feed throughout the mine life.

 

The Project features a comprehensive and integrated infrastructure plan that includes new access roads, power supply systems, water management facilities, a process plant, and storage facilities for tailings and waste rock. Existing support infrastructure will be leveraged, while new installations will address essential gaps in utility access, safety, and environmental control.

 

The total Life-of-Mine capital cost is estimated at $ 416.9 million, comprising:

 

·Pre-production capital of $ 263.6 million (23-month period),

 

·Sustaining capital of $136.2 million (over 17 years), and

 

·Closure capital of $17.2 million (over the last 3 years of the LoM).

 

The cost estimate is a Class 5 estimate (±30%/±50 %) with a 12% contingency, excluding working capital, VAT, escalation, and financing. It is based on budgetary quotes, benchmarks from Latin American projects, and internal cost databases.

 

Operating costs were derived using first principles and local benchmarks. Processing, site services, and general and administrative (G&A) costs were carefully broken down, including labor, power, consumables, and maintenance.

 

1.20.1Risks

 

The most significant potential risks associated with the Project include the hot water management that will be encountered during the mine dewatering effort and socio-political

 

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resistance to the development of the planned mine in Guatemala. The latter is a common risk to most mining projects. It can be mitigated, at least to some degree, with adequate planning and proactive management. The risk associated with water management is not entirely unknown, given the presence of existing dewatering wells and the continued dewatering, treatment, and discharge of underground water.

 

It is important to note that the current mine plan is based on a resource model composed exclusively of Indicated and Inferred Resources, and Inferred Resources are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves. As such, there is a significant degree of uncertainty associated with the tonnages and grades used in the sequencing.

 

The cost of grid power is based on a market survey rather than an actual power supply agreement. A higher power cost would result in increased operating costs.

 

Although the local community is favorable to the development of Era Dorada as an underground mine, there is a potential risk of socio-political opposition to mine development, which could adversely impact the Project development schedule.

 

The ability to achieve the estimated CAPEX and OPEX costs is an important element of the Project’s success. If OPEX increases, then the NSR cut-off would also increase, and all else being equal, the size of the mineable resource would decrease, yielding fewer mineable tonnes.

 

1.21Recommendations

 

·Additional drilling will increase resources and improve understanding and modeling of lithological units.

 

·Definition drilling ahead of blasting will improve the definition of grade boundaries between high-grade veins and low-grade disseminated mineralized material and help minimize unplanned dilution.

 

·A review of mineral resource classification and grade distributions is prudent to ensure accuracy and certainty.

 

·For geotechnical purposes, it is available to characterize and model the geotechnical parameters as domains and placement into the estimation block model.

 

·A comprehensive brownfields exploration program along trend of the main deposit is recommended to explore for additional gold and silver resources that could potentially extend the project’s life.

 

·Optimization of mine plan.

 

·Implementation of power generation in the cooling of the mine water.

 

·Mining Study detailing mining dilution for both mining methods.

 

·Detailed groundwater and dewatering control along LoM.

 

·Develop a detailed mining operating plan that respects all mining activities, accounting for project restrictions, equipment productivity, and limitations.

 

·Complete detailed engineering for the site infrastructure, ensuring optimization of costs, constructability, and operational integration.

 

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·Submit final permitting documentation and ensure all facilities are compliant with local, national, and international environmental standards and regulations.

 

·Continue detailed geochemical testing for waste rock and tailings to confirm long-term environmental stability and support final facility design.

 

·Develop a phased construction plan with critical path scheduling and initiate procurement of long-lead equipment and materials.

 

·Maintain proactive communication with local communities and stakeholders to support social license and minimize construction-related disruptions.

 

·Implement a robust risk mitigation plan for infrastructure development, including contingency planning for stormwater events, equipment delays, and logistics challenges.

 

·Refine cost estimates to Class 5 level or higher, incorporating detailed engineering, contractor bids, and updated procurement quotes to improve accuracy and reduce contingency requirements.

 

·Evaluate project economics under different gold price scenarios, inflation rates, and cost escalations to test project resilience and identify key cost drivers.

 

·Optimize the project scheduling to prioritize higher-grade zones during the initial years of operation, thereby enhancing early revenue generation.

 

·Evaluate alternative production scenarios involving variable feed rates throughout the LoM to improve project flexibility and economic performance.

 

·Conduct a PFS or FS study for Mineral Reserve certification considering potential variations in mining methods and/or stope geometry to identify opportunities for improved resource recovery and economic efficiency.

 

·Use the current capital structure and cost estimates to support investment discussions, including potential financing, offtake agreements, or joint venture opportunities.

 

·Establish early procurement strategies, capital budgeting systems, and contract structures that enable cost discipline and reduce construction risk.

 

·Incorporate local tax regimes, VAT recoverability, depreciation schedules, and financing structures to derive a complete economic picture for stakeholders.

 

·Ensure that projected expenditures for G&A, environmental compliance, and social responsibility are transparently communicated and aligned with local expectations.

 

·Undertake a comprehensive technical and economic evaluation of the dewatering system to identify opportunities for cost reduction and efficiency improvements.

 

·Evaluate alternative technologies, energy-saving strategies, and hydrological modeling to minimize the operational impact of dewatering on OPEX.

 

·Implement a continuous monitoring strategy for gold price fluctuations, with regular updates to the economic model to assess impacts on Net Present Value (NPV), Internal Rate of Return (IRR), and payback period.

 

·Perform updated sensitivity analyses at key decision points to evaluate the Project's resilience under various pricing scenarios.

 

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2INTRODUCTION

 

In January 2025, Aura Minerals Inc (“Aura” or “the Company”) completed the acquisition of the Era Dorada Gold Project—formerly named “Cerro Blanco Project”—and Mita Geothermal Project, located in Jutiapa, Guatemala, near the town of Asunción Mita and the border with El Salvador. The Era Dorada Project (“Era Dorada” or “the Project”) is 100% beneficially owned by Aura. Aura is a public, TSX-listed company trading under the symbol “ORA”, with its head office located at 78 SW 7th St., Miami, FL 33130, USA.

 

Aura commissioned GE21 Consultoria Mineral Ltda. (GE21) to prepare a Technical Report Summary (TRS) for the Project. The Mita Geothermal Project is not considered in this report.

 

This Technical Report Summary titled “Era Dorada Gold Project – Initial Assessment” presents information in compliance with United States Securities and Exchange Commission’s (SEC) Modernized Property Disclosure Requirements for Mining Registrants as described in Subpart 229.1300 of Regulation S-K, Disclosure by Registrants Engaged in Mining Operations (S-K 1300) and Item 601 (b)(96) Technical Report Summary.

 

The purpose of this study is to document a Mineral Resource Estimate, mine design, and preliminary economics of the Project. Bluestone Resources Inc (Bluestone) had previously completed FS studies on the Project in 2019 and updated the results in 2022. These previous studies did not comply with S-K 1300 guidelines; however, in preparation for the Initial Assessment, any relevant information was reviewed and reused where deemed appropriate by the QPs.

 

This report summarizes the work carried out by several consultants, and the scope of work for each company is listed below. Combined, these make up the total Project scope.

 

GE21 scope of work included:

 

·Compiling the technical report, including information provided by other consulting companies;

 

·Establishing an economic framework for the Initial Assessment;

 

·Mine engineering, design, and scheduling;

 

·Designing required site infrastructure;

 

·Estimating mining, process plant, and G&A OPEX and CAPEX for the Project;

 

·Preparing a financial model and conducting an economic evaluation, including sensitivity;

 

·Interpreting the results and making conclusions that lead to recommendations to improve Project value and reduce risks.

 

Kirkham Geosystems Ltd. (Kirkham) scope of work included:

 

·Mineral Resource Estimate.

 

HDJ’s scope of work included:

 

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·Compiling the technical report, including information provided by other consulting companies;

 

·Developing a conceptual flowsheet, specifications, and selection of process equipment;

 

·Designing required plant facilities and other ancillary facilities;

 

·Establishing gold and silver recovery values for doré production on-site;

 

·Evaluating the status of current permits.

 

2.1Qualified Persons

 

The Qualified Persons (QPs) responsible for this independent TRS are Mr. Porfirio Cabaleiro Rodriguez, Homero Delboni Jr., and Garth Kirkham (Table 2-1). Neither GE21 nor the authors of this TRS have had any material interest invested in Aura or any of its related entities.

 

Mr. Porfirio Cabaleiro Rodriguez, Director of GE21 Consultoria Mineral, is a mining engineer, a Fellow of the AIG (FAIG #3708), and has more than 40 years of experience in Mineral Resource and Mineral Reserve estimation. Mr. Rodriguez has sufficient experience relevant to the styles of mineralization and types of deposits under consideration to be considered a QP, as defined by S-K 1300. He is responsible for supervising all sections in this independent TRS, is individually responsible for Sections 2, 3, 4, 5, 13, 15, 16, 18, 19, 24, and 25, and is co-responsible with other QPs for Sections 1, 22 and 23.

 

Dr. Homero Delboni Jr. is a Mining Engineer and Minerals Processing, Ph.D in Minerals Processing and Chartered Professional (Metallurgy) of the Australasian Institute of Mining and Metallurgy (AusIMM #112813), and has more than 40 years of experience in mineral processing. Mr. Delboni has sufficient and relevant experience in mineral processing industrial circuits to be considered a QP as defined by S-K 1300. He is individually responsible for Sections 10, 14, and 17 and is co-responsible with other QPs for Sections 1, 22, and 23.

 

Mr. Garth Kirkham, P.Geo. and Principal of Kirkham Geosystems Ltd., is a geophysicist and geologist (EGBC #30043) and has more than 30 years of experience in experience supplying 3D geoscience modeling, geological and geophysical consulting services to the mining, environmental and oil & gas industries. He has sufficient and relevant experience in mineral deposit geology and resource estimation to be considered a QP as defined by S-K 1300. He is individually responsible for Sections 6, 7, 8, 9, 11, 20, and 21 and is co-responsible, with other QPs, for Sections 1, 22 and 23.

 

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Table 2-1: List of QPs and related responsibilities

 

QP Section Responsibility Site Visit Responsibility
Porfirio Cabaleiro Rodriguez, FAIG 2, 3, 4, 5, 13, 15, 16, 18, 19, 24 and 25 and partially 1, 22 and 23   Author and Peer Review
Dr. Homero Delboni Jr., AusIMM 10, 14 and 17 and partially 1, 22 and 23   Author and Peer Review
Garth Kirkham, P. Geo. 6, 7, 8, 9, 11, 20 and 21 and partially 1, 22 and 23 May 8, 2017; Sep 21-22, 2017; Apr 24-28, 2018; Feb 16-22, 2020, Jan 10-15, 2021 Author and Peer Review

 

Source: GE21, 2025.

 

2.2Site Visits and Scope of Personal Inspection

 

Garth Kirkham, P. Geo., first visited the property on May 8, 2017, to satisfy the site visit requirements related to the 2017 Technical Report. The site visit included an inspection of the property, offices, underground vein exposures, core storage facilities, the water treatment plant, stockpiles, and a tour of major centers and surrounding villages that are most likely to be affected by any potential mining operation.

 

Since 2017, Mr. Kirkham has visited the property numerous times for extended periods to develop and implement data-gathering and sampling methods and procedures. He also worked with Bluestone geologists to develop drill programs and supervise the interpretation and wireframe modeling, in addition to vetting and reviewing QA/QC procedures. On September 21-22, 2017, Mr. Kirkham inspected the progress of the recommended historic drill core rehabilitation program and initiated the structural studies. From April 24 to 28, 2018, the site visit focused on advancing the planning and development of sampling and drilling, as well as supporting lithological and structural modeling. From February 16 to 22, 2020, Mr. Kirkham assisted with the planning and development of advanced drilling and sampling. He provided guidance on lithology and high-grade vein modeling for resource estimation. From January 10 to 15, 2021, Mr. Kirkham validated drill and sample data, refined high-grade models, reviewed low-grade models, and provided guidance for finalizing an open pit bulk tonnage resource scenario.

 

The other QPs relied upon the observations of Mr. Kirkham, who visited the site.

 

2.3Effective Date and Sources of Information

 

This report is based on information collected by the QPs during site visits and on additional information provided by Aura throughout the course of GE21’s analysis. Other information was obtained from the public domain. GE21 has no reason to doubt the reliability of the information provided by Aura.

 

Aura and its consultants provided GE21 with the information that was used to prepare this TRS, specifically during the execution of the work that is described herein. This work reflects the technical and economic conditions at the time that it was executed. The authors, whenever possible, executed independent verification of the data they received, in addition to conducting field visits to corroborate the data. This information was supplied in the form of an exploratory

 

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drilling database, certifications, maps, technical reports, and a topographical survey. The data is a combination of historical and newly generated information.

 

The results, images, and illustrations presented in this TRS have been generated from information provided and compiled by Aura through data organized in spreadsheets, internal and third-party technical reports, and supplemental information obtained from the Aura technical team. Exceptions will be subtitled for the source reference.

 

The effective date for this Initial Assessment is December 31, 2024, related to the gold price considered for the Project. The authors believe that no relevant data concerning the Mineral Resources Estimate were produced after this date.

 

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3PROPERTY DESCRIPTION

 

3.1Property Location

 

The Project is located in southeast Guatemala, in the Department of Jutiapa, approximately 160 km by road from the capital, Guatemala City (Figure 3-1), and approximately 9 km west of the border with El Salvador. The nearest town to the Project is Asunción Mita, a community of approximately 18,500 people situated approximately 7 km west of the Project. The exploitation license covers 15.25 km2 and lies entirely in the municipality of Asunción Mita.

 

 

 

Figure 3-1: Project location map

 

Source: Bluestone, 2021.

 

The location of the mineral resources relative to the property boundary is shown in Figure 3-2.

 

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Figure 3-2: Location of Mineral Resources relative to property boundary

 

Source: Bluestone, 2022.

 

3.2Property Description and Tenure

 

The coordinates of the 15.25 km2 exploitation license are recorded in Decree DIC-CM-158-05 and are shown in Figure 3-3. The perimeter of the area is described as having the UTM X and Y coordinates shown in Figure 3-3 and Table 3-1.

 

Table 3-1: Coordinates of exploitation license “Era Dorada”

 

Latitude Longitude
210500 1589500
213000 1589500
213000 1589000
214000 1589000
214000 1585000
210500 1585000

 

Source: Bluestone, 2019.

 

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Figure 3-3: Era Dorada exploitation license coordinates

 

Source: Bluestone, 2019.

 

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3.3Royalties

 

The Project is subject to two royalties, both of which have been included in the economic analysis and cash flow model. Table 3-2 outlines the assumed royalty terms.

 

Table 3-2: Royalty assumptions

 

Parameter Unit Value
Guatemalan Government Royalty % NSR 1.00
Third-Party Royalty % NSR 1.05*

 

Note: *1.05% royalty has been grossed up to account for country withholding tax.

 

Source: Bluestone, 2021.

 

3.4Environmental

 

The Project is following Guatemala environmental laws and regulations and has all necessary permits to proceed with developing the underground mine and construction of the process facilities, subject to future operations adhering to the conditions of the existing permits.

 

However, the Project design has changed since 2007 and requires permit amendments. Additionally, new baseline studies (EIA) and permits are necessary for infrastructure components such as power lines.

 

The current permits and permit amendments are presented in Section 17.

 

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4ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE, AND PHYSIOGRAPHY

 

4.1Access

 

Current road access to the site is via the Pan-American Highway (Highway CA1) through the town of Asunción Mita. Existing infrastructure is in place to provide year-round access to the site. The topography is relatively flat, with rolling hills.

 

Guatemala has 400 km of coastline and claims its territorial waters extend 22 km outward, plus an exclusive economic zone of 370 km offshore. Hurricanes and tropical storms sometimes affect coastal regions.

 

The five main ports in Guatemala and their main activities are listed below:

 

·Atlantic Ports:

 

oPuerto Santo Tomás de Castilla (containers);

 

oPuerto Barrios (containers).

 

·Pacific Ports:

 

oPuerto San José (liquids);

 

oPuerto Quetzal (multi-use);

 

oPuerto Champerico (fishing).

 

Puerto Santo Tomás de Castilla is the most important port on the Atlantic coast of Guatemala. This cargo terminal can handle a variety of cargo (e.g., containers and roll-on, roll-off (RoRo)), as well as general and liquid bulk cargo, passenger ships, vehicle carriers, and barges. The port facilities are approximately 290 km northeast of Guatemala City. The total distance from Santo Tomás de Castilla to the Project site is approximately 440 km.

 

Puerto Quetzal, which is the most important port on the Pacific Coast, has the most modern installations. It is mainly a dry bulk cargo terminal; however, it also handles containers, RoRo, general bulk cargo, and liquid bulk cargo. The port facilities are about 100 km South of Guatemala City. The distance from Puerto Quetzal to the Project site using the coastal highway is approximately 310 km. Puerto Quetzal is 2,050 nautical miles from Los Angeles.

 

These two ports handle nearly 80% of the sea traffic to Guatemala. Guatemala’s Empresa Nacional Portuaria is a state-owned corporation of the Guatemalan port facilities.

 

The nearest airport to attend the region of the Project is Aurora International Airport, in Guatemala City, and 114 km far from the Project, in a 2-h travel by car. It offers a large range of flights and international connections.

 

4.2Climate

 

The climate and vegetation at the Project site are typical of a tropical dry forest environment. The wet season is typically from May to October. The average annual rainfall is

 

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1,350 mm. Daily highs reach 41°C, and lows reach 10°C. The average annual pan evaporation rate is 2,530 mm, with an annual average humidity of 62%. Classified as Zona Oriental, the principal characteristics of the region are a deficiency of rain for much of the year with high ambient daytime temperatures.

 

4.3Physiography

 

The Project is located on a hill with two peaks. The surrounding areas are relatively flat with minimal undulation. A photo showing the typical landscape around the mine property is included in Figure 4-1.

 

 

 

Figure 4-1: Typical landscape in the Project area, looking South

 

Source: Bluestone, 2022.

 

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Most of the vegetation in the Project area loses its foliage because of a lack of precipitation to support growth during the winter months of November through April.

 

The Project occurs within a south-southwest trending ridge that extends from higher ground to the north, outward into the basin and floodplain deposits of the Rio Ostua. The elevation of the upper part of the ridge is in excess of 600 masl. The elevation of the basin and floodplain deposits is about 460 to 490 masl.

 

The west side of the ridge is flanked by a south-southeast-trending perennial drainage called Rio Tancushapa. The east side of the ridge is flanked by a seasonal drainage called Quebrada El Tempisque, which also trends to the south-southeast. These drainages join to the south-southeast of the Project area and flow into the Rio Ostua about 4 km down gradient.

 

The regional area is generally hilly to mountainous, with broad flood plains formed by some of the larger streams and rivers. Three dormant volcanoes are within sight of the Project area: Ixtepeque to the north, Suchitan to the northwest, and Las Viboras to the southwest.

 

4.4Local Resources & Infrastructure

 

The Project is situated in proximity to a number of communities, the largest one being Asunción Mita, with a population of approximately 18,500 people.

 

There is no record of any previous exploitation in the area; however, with the closure of Goldcorp’s Marlin Mine in late 2017, it is anticipated that a significant contingent of Guatemalan-trained labor will be available for employment at Era Dorada. As such, the Project intends to hire the majority of operations staff locally and has allowed for the cost of training programs within the Owner’s budget.

 

The local mine workforce is expected to live in the surrounding communities and provide their own transportation to and from the mine site due to the proximity of the population centers relative to the Project site (Figure 4-2). Employees from distant areas further than Jutiapa and expatriate employees will be housed in the on-site camp.

 

La Barranca power substation is located south of Asunción Mita, approximately 10 kilometers to the west of the Project. The substation has a capacity to supply up to 20 MW of power.

 

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Figure 4-2: Population centers near the Project area

 

Source: Bluestone, 2022.

 

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5HISTORY

 

There is no evidence of exploration activity on the Era Dorada property (formerly “Cerro Blanco”) before 1997. Mar-West Resources Ltd. (Mar-West), a Canadian exploration company, had been working in adjacent Honduras since 1995 and expanded their gold prospecting activities into southern Guatemala in 1997. The Cerro Blanco property was identified by Mar-West by sampling densely silicified boulders, in some cases cut by chalcedonic veinlets, during an initial reconnaissance evaluation of an area known for active hot springs. Traverses over the hill at Cerro Blanco yielded surface rock assays of 1 to 3 g Au/t. An exploration concession was subsequently applied for and granted in late 1997. Mar-West drilled nine reverse circulation (RC) holes from April to June 1998, which tested near-surface potential to shallow depths of 100 to 150 m. At least seven holes contained one or more intercepts of 5 to 15 m grading 1 to 5 g Au/t, with the occasional 10 to 20 g Au/t interval, and were sufficient to justify continued exploration on the property.

 

In October 1998, Mar-West’s holdings in Honduras and Guatemala were purchased by Glamis Gold Ltd. (Glamis) primarily to acquire the San Martin deposit in Honduras. Mar-West geologists continued to manage the Cerro Blanco exploration program through March 1999. The sinter area was soil sampled and trenched, and drilling was advanced to hole 19 when geophysical orientation surveys were undertaken. A further 331 drill holes were completed up until 2006.

 

Goldcorp became the sole proprietor of the Cerro Blanco Gold Project through the purchase of Glamis in November 2006. Goldcorp undertook a comprehensive exploration program from 2006 to 2012, including additional surface exploration, over 3.4 km of underground development, and 43,016 m of surface and underground drilling. Exploration activities at the Cerro Blanco property by Goldcorp included the following:

 

·Surface soil geochemistry;

 

·Surface rock geochemistry;

 

·Surface geological mapping;

 

·Construction of the north and south ramp access and ventilation raises;

 

·Underground geological mapping;

 

·Underground chip sampling;

 

·Surface and underground diamond drilling.

 

Several unpublished feasibility studies were completed by Goldcorp from 2011 to 2014. Kappes, Cassidy & Associates (KCA) and Golder Associates (Golder) completed an FS for the Project in May 2012. After this initial FS, Goldcorp issued a new geological model and requested KCA and Golder to update the FS in 2013 using a revised mine design, mine development, mine operation, and capital costs. In 2014, an internally updated FS was produced with optimized mine stope parameters and the mine schedule and costing information that was updated by Maptek.

 

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On January 4, 2017, Bluestone entered into an agreement with Goldcorp Inc. (Goldcorp) to acquire 100% of Minerales Entre Mares de Guatemala, S.A. (Entre Mares, or EM), which was Goldcorp's indirect wholly owned Guatemalan subsidiary which holds a 100% interest in Cerro Blanco. On successful closure of the deal, Entre Mares became a wholly owned subsidiary of Bluestone, a Canadian company headquartered in Vancouver, British Columbia.

 

In January 2025, Aura completed the acquisition of the Cerro Blanco Project and Mita Geothermal Project, located in Jutiapa, Guatemala, near the town of Asunción Mita and the border with El Salvador. The Cerro Blanco Project is 100% beneficially owned by Aura.

 

5.1Data Validation History

 

Historical core logging, sampling, and QA/QC procedures were reviewed by Golder in 2014. Ten core samples were collected from quarter-sawn NQ core, and selected drill hole collars were surveyed using a GPS. Assayed gold and silver grades were found to be consistent with those reported by Goldcorp. Golder was satisfied that the drill hole data was collected in a manner consistent with industry best practice standards.

 

As part of the core logging data verification, Golder compared a selection of core logs against half-core stored at the Project site. Five half-core drill holes were reviewed from the North and South deposits. The Excel files were reviewed first, and drill holes were selected that represented the typical mineralization style for each deposit. In addition, 10 verification samples were taken from these drill holes. Each verification sample was a half-core sample sawed in half again, with the quarter sample sent for analysis and the other quarter returned to the core racks. Table 5-1 summarizes the samples selected for core logging review and verification sampling.

 

Table 5-1: Verification samples

 

Drill Hole ID

Duplicate

Sample No.

Original

Sample No

From (m) To (m) Deposit

Metal

Analyzed

Rock Type
CB-152 205873 82225 128 129 North Au, Ag Lapilli Tuff
CB-152 205874 82226 129 130 North Au, Ag Lapilli Tuff
CB-200 205884 407101 156 157 South Au, Ag Quartz Tuff
CB-200 205885 407102 157 158 South Au, Ag Quartz Tuff
CB-241 205891 404849 111.4 112.6 South Au, Ag Conglomerate
CB-241 205892 404850 112.6 113.5 South Au, Ag Fault
CB-254 205895 414397 100.5 102 South Au, Ag Volcaniclastic sediments
CB-254 205896 414398 102 103.5 South Au, Ag Volcaniclastic sediments
CB-10-15 205871 435941 135 136.23 North Au, Ag Lapilli Tuff
CB-10-15 205872 435943 136.23 137.46 North Au, Ag Lapilli Tuff

 

Source: Goldcorp, 2014.

 

Samples were sawed and bagged under Golder's supervision and were transported off-site via helicopter and plane to Canada and then by ground transportation to ALS Chemex laboratories in Sudbury for sample preparation and analysis.

 

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A comparison of the Excel files against the drill core indicated an excellent match between the core logs and the retained core.

 

Table 5-2 is a list of the drill hole collar surveys completed by Golder.

 

Table 5-2: Drill hole collar survey (NAD 27 Zone 16N)

 

Drill Hole ID Golder Cerro Blanco
Easting Northing Easting Northing
C 10 08 212015.1 1587867 212009 1587748
C 11 12 211906.8 1587714 211904 1587605
C 11 15 211969.7 1587769 211966 1587655
C 11 18 211866.4 1587405 211873.2 1587297
C 11 21 211901.6 1587414 211898.9 1587307
C 151 212025.1 1587821 212020.8 1587707
C 247 211985.5 1587315 211978.8 1587202

 

Source: Goldcorp, 2014.

 

Eight drill sites were visited, with multiple drill holes located at some sites. Casings had been removed for most drill holes. The data collected was a mixture of pre-Goldcorp drill holes (2006 or earlier) and drilling completed by Goldcorp during 2010 and 2011. All drill holes from the surface were grouted to prevent water flow into the underground workings.

 

Approximately 5% of the drill holes (20 holes) were subjected to data verification checks by Golder. The 20 selected holes, summarized in Table 5-3, included a variety of historical data as well as some of the more recent holes. The data verification checks consisted of the following:

 

·comparison of final assays to the original laboratory certificates;

 

·analysis of external laboratory duplicate assays by generating XY scatter plots;

 

·review of downhole survey measurements to identify anomalous changes to hole orientation.

 

Table 5-3: Drill holes selected for data verification

 

Drill Hole ID
CB-012 CB-200
CB-016 CB-227
CB-063 CB-244
CB-078 CB-247
CB-095 CB-305
CB-10-02 CB-309
CB-120 CB-314
CB-142 CB-345
CB-146 CB-357
CB-151 CB-362

 

Source: Goldcorp, 2014.

 

For the 20 holes reviewed, the comparison of final assays to the original assay certificates did not identify any material differences in assay values.

 

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External laboratory duplicate assays were reviewed to assess the reliability of the primary assay laboratory. XY scatter plots were generated for each of the 20 holes. With the exception of a few outliers, the majority of the data compared well. Figure 5-1 illustrates an example of the XY scatter plots used to compare assay results.

  

 

 

Figure 5-1: Example of XY scatter plot for hole CB34

 

Source: Goldcorp, 2014.

 

5.2Historic Resources

 

Indicated and Inferred Resources were initially reported in 2008 for the Cerro Blanco Project at an 8.0 g/t gold equivalent cut-off grade as follows in Table 5-4.

 

Table 5-4: Indicated and Inferred Mineral Resource Estimate (2008)

 

Resource

Type¹

Tonnes

(kt)

Gold Grade

(g/t)

Contained Gold

(koz)

Silver Grade

(g/t)

Contained

Silver (koz)

Contained

Equivalent

Ounces of Gold²

Indicated 2,500 15.65 1,266 72 5,826 1,324
Inferred 1,400 15.3 665 59.6 2,589 691

 

Notes:

1.The Mineral Resources have been calculated in accordance with definitions adopted by the Canadian Institute of Mining, Metallurgy, and Petroleum on August 20, 2000. Employees of Glamis Gold Ltd., under the supervision of James S. Voorhees, Executive Vice President of Operations and Chief Operating Officer, have prepared these calculations.

2.The conversion of silver ounces to gold-equivalent ounces is at a ratio of 100 silver ounces to one gold-equivalent ounce.

Source: Voorhees, 2008.

 

Subsequently, Mineral Resources were reported as in-situ resources at cut-off grades of 1.0 g/t Au and 4.0 g/t Au in 2014. Blocks were classified based on drill spacing (sample distances) and the number of drill

 

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holes. In-situ Mineral Resources are summarized in Table 5-5. No wireframing was performed around these blocks.

 

Table 5-5: In-situ Mineral Resources (2014)

 

In-situ Mineral Resources (veins)
COG (g/t) Mt Au g/t Ag g/t Au Moz Ag Moz
Indicated In-situ* Mineral Resources
1 19 3.38 13.9 2.06 8.5
4 3.85 9.71 35.2 1.2 4.4
Inferred In-situ* Mineral Resources
1 2.3 3.1 8.5 0.22 0.6
4 0.33 10.8 19.8 0.11 0.2

 

Note: *Reported in-situ Mineral Resources do not consider mineral availability by underground or open pit mining methods.

Source: Goldcorp, 2014.

 

In 2016, Goldcorp listed resources for Cerro Blanco within its Annual Report. The last public statements by Goldcorp outlined historical resources of 2.05 Mt grading 12.69 g/t for 840,000 oz of gold in the Indicated category, as well as 0.75 Mt grading 9.34 g/t for 230,000 oz of gold in the Inferred category.

 

The Indicated and Inferred resources are historical estimates and use the categories set out in NI 43-101. These resources are effective as of June 30, 2016, and are disclosed in Goldcorp’s press release dated October 26, 2016. Resources were estimated using US$ 1,400/oz AU and US$ 20/oz Ag. Given the source of the estimates, Bluestone considers them reliable and relevant for the further development of the Project; however, a qualified person has not done sufficient work to classify the historical estimates as current Mineral Resources or Mineral Reserves, and the Company is not treating the historical estimates as current Mineral Resources or Mineral Reserves.

 

In 2021, the mineral resource estimate was based on a scenario that considered open-pit mining methods and is reported at a base case above a 0.4 g Au/t cut-off, as tabulated in Table 5-6.

 

Mineralized material from mining activities conducted up to 2021—including ramp development and access—was stockpiled on-site and segregated for future processing. Correlograms for gold and silver were created and employed to estimate the stockpile resources using ordinary kriging. The estimate was validated using nearest neighbor and inverse distance methods.

 

In addition to the open-pit resources, high-grade vein material located below the pit shell remains a potential target for limited underground mining meth.

 

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Table 5-6: Mineral Resource statement (2021)

 

Resource Category Tonnes (kt) Au Grade (g/t) Ag Grade (g/t) Contained Gold (koz) Contained Silver (koz)
Measured 40,947 1.8 7.9 2,382 10,387
Indicated 22,595 1.0 4.2 706 3,058
Measured & Indicated 63,542 1.5 6.6 3,089 13,445
Inferred 1,672 0.6 2.1 31 112
           
Below Pit (Indicated)* 189 5.7 13.4 35 82
Stockpile (Measured) 30 5.4 22.6 5 22

 

Notes: The mineral resource statement is subject to the following:

1.All mineral resources have been estimated in accordance with Canadian Institute of Mining and Metallurgy and Petroleum (CIM) definitions, as required under National Instrument 43-101 (NI 43-101), with an effective date of December 31, 2020.

2.Mineral resources reported demonstrate a reasonable prospect of eventual economic extraction, as required under NI 43-101; mineral resources are not mineral reserves and do not have demonstrated economic viability.

3.*Underground mineral resources are reported at a cut-off grade of 3.5 g Au/t. Cut-off grades are based on a price of US$ 1,600/oz gold, US$ 20/oz silver, and a number of operating cost and recovery assumptions, plus a contingency.

4.Numbers are rounded.

5.The mineral resources may be affected by subsequent assessment of mining, environmental, processing, permitting, taxation, socio-economic and other factors.

6.An inferred mineral resource has a lower level of confidence than that applying to an indicated mineral resource and must not be converted to a mineral reserve. It is reasonably expected that the majority of inferred mineral resources could be upgraded to indicated mineral resources with continued exploration.

7.Mineral Resources are inclusive of mineral reserves.

Source: Kirkham, 2021.

 

5.3Historic Reserves

 

The mining stope and sub-level designs with external, backfill, and planned dilution, along with the mining recovery factor applied, determined the 2019 Mineral Reserve estimate shown in Table 5-7.

  

Table 5-7: Mineral Reserve estimate (2019)

 

Class

Diluted Tonnes

(kt)

Au Grade (g/t) Ag Grade (g/t) Au Onces (koz) Ag Ounces (koz)
Proven 313 8.3 31.4 83 315
Probable 3,131 8.5 32.3 857 3,256
Total 3,444 8.5 32.2 940 3,571
   
Notes:

1.The Qualified Person for the Mineral Reserve estimate is Michael Makarenko, P. Eng., of JDS Energy & Mining Inc.

2.Effective date: January 29, 2019. All Mineral Reserves have been estimated in accordance with Canadian Institute of Mining and Metallurgy and Petroleum (CIM) definitions, as required under NI 43-101.

3.Mineral Reserves were estimated using a $ 1,250 /oz gold price and a gold cut-off grade of 3.5 g/t. Other costs and factors used for gold cut-off grade determination were mining, process, and other costs of $ 109.04/t, transport and treatment charges of $ 5.00/oz Au, a royalty of $ 24.84 /oz Au, and a gold metallurgical recovery of 95%.

4.Silver was not used in the estimation of cut-off grades but is recovered and contributes to the project cash flow.

5.Tonnages are rounded to the nearest 1,000 t; metal grades are rounded to one decimal place. Tonnage and grade measurements are in metric units; contained gold and silver are reported as thousands of troy ounces.

6.Rounding, as required by reporting guidelines, may result in summation differences.

Source: Bluestone, 2019.

 

In 2021, Mineral reserves for the Cerro Blanco Gold Project were estimated at 53.9 Mt at an average grade of 1.64 g/t of gold for 2,846,000 ozs and 7.27 g/t of silver for 12,602,000 ozs, as summarized in Table 5-8. The Mineral Reserve Estimate (MRE) was prepared by G Mining Services Inc. (GMS)

 

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Table 5-8: Cerro Blanco Gold Project open pit Mineral Reserve estimate (2022)

 

Reserve Category Tonnage (kt) Gold (g/t) Gold (koz) Silver (g/t) Silver (koz)
Proven 37,618 1.89 2,286 8.33 10,084
Probable 16,279 1.07 560 4.81 2,518
Proven & Probable 53,896 1.64 2,846 7.27 12,602
   
Notes:

1.CIM definitions were followed for mineral reserves.

2.The effective date of the estimate is Nov 1, 2021.

3.Mineral reserves are estimated at a cut-off grade of 0.50 g/t Au Eq.

4.Mineral Reserves are estimated using the following long-term metal prices (Au = US$ 1,550/oz and Ag = US$ 20/oz).

5.The bulk density of ore is variable but averages 2.70 t/m3.

6.The average strip ratio is 2.7:1.

7.The average mining dilution factor is 6.7%.

8.Other costs and factors used for gold cut-off grade determination were process, G&A, and other costs of $ 21.17/t, a royalty of $ 31.60 /oz Au, and gold and silver metallurgical recoveries of 91% and 85%, respectively.

9.Tonnages are rounded to the nearest 1,000 t; metal grades are rounded to two decimal places. Tonnage and grade measurements are in metric units; contained gold and silver are reported as thousands of troy ounces.

10.The mineral reserves may be affected by subsequent assessment of mining, environmental, processing, permitting, taxation, and socio-economic factors.

11.Mineral resources are inclusive of mineral reserves.

Source: GMS, 2021.

 

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6GEOLOGICAL SETTING, MINERALIZATION, AND DEPOSIT

 

6.1Introduction

 

The geology of Guatemala comprises rocks that are divided into two tectonic terrains due to the collision between the North American, Caribbean, and Cocos tectonic plates during the Upper Cretaceous, 70 to 90 million years ago. The Maya Block to the north is characterized by igneous and metamorphic basement rocks overlain by late Palaeozoic metasediments. Mesozoic red beds, evaporites, and marine limestones overlie these rocks, and a karst landscape formed in the thick limestone units across the north of the country. By contrast, southern Guatemala, south of the Motagua Valley, belongs to the Chortis Block, representing the northern part of the Caribbean Plate. This region forms an active volcanic arc termed the Central American Volcanic Arc (Figure 6-1), which continues from the Guatemala-Mexico border along the Pacific side of Central America into central Costa Rica, with most of the major eruptive events having occurred in the Tertiary and Quaternary.

 

 

 

Figure 6-1: Location of Era Dorada and other deposits in the Central American Volcanic

 

Source: Bluestone, 2020.

 

6.2Regional Geology of Southern Guatemala

 

Southern Guatemala, El Salvador, Honduras, and Nicaragua are located within the Chortis continental crustal block. The tectonic event that sutured the Chortis block to the North American craton took place between 66 and 70 million years ago along the east-west-striking Polochic-Montagua fault system that crosses

 

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southern Guatemala (Figure 6-2Figure 6-2). Three regional east-west trending, left-lateral transform faults form the plate collision boundary, defined by the Polochic, Motagua, and Jocotan fault systems from north to south. Nearer the Cerro Blanco deposit, other major regional structures that strike north-northeast, such as the Jalpatagua and Ipala faults, are important local structures.

 

A large group of granitic stocks and batholiths intruded the suture zone south of the Polochic-Montagua fault with ages of 35 to 85 million years. These broadly brackets, both temporally and spatially, the collision event (Donnelly et al., 1990).

 

 

 

Figure 6-2: Regional structural map of Guatemala

 

Source: Bluestone, 2021.

 

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The Jocotan Fault is generally considered the southernmost major suture-related fault. It is an east-west fault with considerable Late Cretaceous dip-slip movement (south side down), but it had little or no Tertiary transcurrent movement. Era Dorada is located about 50 km south of the Jocotan Fault.

 

The ancestral Middle America Trench developed at this time. The Pacific Oceanic plate is subducted beneath Central America and is the principal driving force for volcanic and intrusive igneous activity throughout Central America along this boundary trench. The earliest documented volcanic outpouring on the Chortis block was the Paleocene (about 55 to 65 million years ago) (Pindell and Barrett, 1990).

 

In Costa Rica and Panama, a series of west-northwest-trending (arc-parallel) back-arc basins developed. These accumulated tuffaceous sediments continuously from the Eocene (about 55 million years) to the present (Donnelly et al., 1990). The principal periods of Andean-style calc-alkaline volcanism in the Chortis block include the Paleocene-Eocene (relatively minor), Oligocene (major), and Miocene-Pliocene (the biggest) (Pindell and Barrett, 1990).

 

The Polochic-Montagua suture was reactivated as a sinistral (left-lateral) transform fault that displaced the Chortis block 130 km eastward with respect to the North American craton. Movement took place from 6 to 10 million years ago (Deaton and Burkart, 1984). An associated extension was accommodated by a series of north-south grabens across southern Guatemala and western Honduras. Back-arc rift basins developed adjacent to northwest-striking normal faults all along western Central America. The Nicaraguan Rift began to form about 7 million years ago and continues to subside today. Bimodal, rhyolite-basalt volcanism began during this event and, by 7 million years ago, was widespread throughout the western half of the Chortis block.

 

A large number of Central American gold deposits, including Marlin and Era Dorada, occur within a narrow belt parallel to the western Central American coast from southern Guatemala through to Panama. The geology of Guatemala comprises rocks that are divided into two tectonic terrains due to the collision between the North American, Caribbean, and Cocos tectonic plates during the Upper Cretaceous, 70 to 90 million years ago. The Maya Block to the north is characterized by igneous and metamorphic basement rocks overlain by late Palaeozoic metasediments. Mesozoic red beds, evaporites, and marine limestones overlie these rocks, and a karst landscape formed in the thick limestone units across the north of the country. By contrast, southern Guatemala, south of the Motagua Valley, belongs to the Chortis Block, representing the northern part of the Caribbean Plate. This region forms an active volcanic arc termed the Central American Volcanic Arc (Figure 6-1), which continues from the Guatemala-Mexico border along the Pacific side of Central America into central Costa Rica, with most of the major eruptive events having occurred in the Tertiary and Quaternary.

 

This metallogenic belt follows the volcanic arc, and precious metal deposits are clearly related in space and time to Miocene-Pliocene extensional tectonics and associated bimodal basalt-rhyolite volcanism. Published age dates cluster between 4 and 8 million years. Argon-argon dating (40Ar-39Ar) of vein adularia from Era Dorada returned a date of 4.93 ± 0.47 Ma.

 

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6.3Local Geology

 

The Project deposit is a classic hot springs-related, low-sulfidation quartz-chalcedony-adularia-calcite vein system. It was localized along a structural corridor created during the late Miocene- Pliocene tectonic extension within the active Central American Volcanic Arc. Deep penetrating faults and local bimodal igneous activity drove the Cerro Blanco hydrothermal system and the formation of the gold deposit.

 

The Project lies within the volcanic province, with the principal rock units being Tertiary volcanic, volcaniclastics, and sediments, including ignimbrites, siltstone, limestones, and conglomerates, that are intruded by andesitic and rhyolitic dykes. Recent basalt lava flows form the youngest rocks in the area in addition to locally derived volcanic sediments.

 

The gold- and silver-bearing veins and upper unit of silicified sediments (Salinas unit) occupy a north-trending graben bounded by a fault (termed the East Fault), representing a major structural feature that separates the main Era Dorada gold deposit from the Mita geothermal field immediately to the east.

 

To the north, the graben is concealed beneath Quaternary basalt flows, and to the south, it is concealed by recent alluvium. Rhyolite/dacite domes underlie the extreme northeast portion of the district. Active hot springs occur immediately south of Cerro Blanco hill.

 

Figure 6-3 shows a simplified geological map for Cerro Blanco.

 

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Figure 6-3: Geological map of Era Dorada

 

Source: Pratt and Gordon, 2019.

 

6.3.1Lithology

 

The oldest rocks at Cerro Blanco Gold Project, intersected in deep drill holes, belong to the Mita Group (Pliocene-Miocene). This group exhibits a great variety of volcanic and sedimentary rocks with important marker beds that are crucial for understanding complex structural geology. Thicknesses seem fairly constant, with little evidence of growth faulting or internal unconformities during their accumulation (Figure 6-4).

 

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Figure 6-4: Lithostratigraphy & lithology at Era Dorada

 

Source: Pratt and Gordon, 2019.

 

The deeper parts of the Mita Group are dominated by volcaniclastic rocks (Mvo, mass flow deposits, conglomerates) with intercalated auto-brecciated and amygdaloidal porphyritic andesites (lithology code PA). There is a distinctive unit of dark grey siltstones and fine sandstones (Silt), frequently with syn- sedimentary disruption. The sequence is capped by a major unit of andesitic-dacitic tuff (Mcv) (Figure 6-5), which erupted in a single event. This is at least 50 m thick and rich in broken crystals and small pumice lapilli. It shows a weak compaction fabric or welding (refer to the photographs in Figure 6-5).

 

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Figure 6-5: Examples of andesitic lapilli tuff (Mcv)

 

Source: Pratt and Gordon, 2019.

 

The tuff is overlain by sandstones (Mss), followed by a nodular micritic to shelly, oyster-rich limestone (Mls, see Figure 6-6), which is the most distinctive rock at Era Dorada. This limestone sequence is about 20 m thick and includes calcarenites (Msc).

 

The limestone is overlain by a thick sequence of relatively massive, brick-red to light grey siltstone and fine sandstone (Mbt). This distinctive rock has local accretionary lapilli, horizons of flaser and ripple cross-bedded fine sandstone, and local calcareous concretions. The Mbt sequence is divided into lower and upper parts by an andesitic crystal tuff (Mat). It is also punctuated by intervals of clean, well-sorted, fine-grained conglomerate (Mss). These can be rich in metamorphic vein quartz pebbles and dark grey schist, indicating a metamorphic hinterland. In the north part of the property, there is a second major package of limestone (Mlm) (Figure 6-3), in turn overlain by further massive siltstones (Mbt).

 

The Mita Group is overlain by the Salinas Group (Svc). This is a complex sequence of interbedded plant-rich siltstones, mudstones, sandstones, conglomerates, mass flow deposits, phreatic breccias, and hot spring sinters. The Salinas unit, of probable Pliocene age, was previously considered to unconformably overlie the Mita unit, which was then assigned to the Eocene-Oligocene. The presence of the unconformity is certainly suggested by the structural culmination defined by the Mita limestone. However, thin sinter horizons are observed interbedded with siltstone at the top of the Mita unit, a situation that requires that the Mita and Salinas

 

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are part of a single, uninterrupted succession. This interpretation implies that the Mita part of the succession was in place before the mineralization commenced, whereas the overlying Salinas part accumulated during the mineralization event (Sillitoe, 2018).

 

 

 

Figure 6-6: Examples of limestones (Mls)

 

Source: Pratt and Gordon, 2019.

 

The syn-mineral Salinas unit is believed to have accumulated progressively in a low-relief graben characterized by a shallow groundwater table. The Salinas conglomerate was presumably derived by erosion of the flanking horst blocks as relief was created during the active faulting. The topographic inversion required to explain the current prominent position of the graben fill is ascribed to the silicic character of the Salinas unit and its consequent resistance to erosion.

 

Where the paleo-groundwater table intersected the paleosurface, siliceous sinter was precipitated—a situation that must have prevailed on several occasions for relatively protracted time intervals to produce the main sinter horizons. The presence of abundant reed casts in the sinter shows that its formation encroached on marshy ground (Figure 6-7). Where the paleo-groundwater table was several meters below the paleosurface, a conglomerate in its immediate vicinity was silicified, and the vadose zone above it was subjected to steam-heated alteration. The steam-heated alteration, containing cristobalite, kaolinite, and possible alunite (an advanced argillic assemblage), was the product of acidic solutions formed by the condensation of ascendant H2S-bearing steam into downward-percolating groundwater. The overall result is an interlayered sequence of sinter, silicified conglomerate, and steam-heated alteration (Figure 6-7).

 

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The Salinas Group is characterized in the mineralized area by widespread chalcedonic alteration, which can make identifications difficult, and elsewhere by strong clay alteration. In some places, rock fragments have concentric chalcedony coats (pisoliths), implying they accumulated in a hot spring pool. Silicified reed fragments are common and locally upright in their original growth position. Rare gastropods were observed.

 

 

 

Figure 6-7: Silicified reed fragments

 

Source: Pratt and Gordon, 2019.

 

The sequence also includes rhyolitic tuffs and a rhyolite cryptodome / flow dome (Rp), both with bipyramidal, embayed quartz crystals. A dacite cryptodome or flow dome (dp) also crops out around the Era Dorada village and is observed in drill holes in the hanging wall of the East Fault (Figure 6-4). It has no quartz crystals but distinctive, isolated, long hornblende phenocrysts. Sediment dykes, common in geothermal districts, where they form the feeders to sand and mud volcanoes, are common in the Salinas Group.

 

A typical log of the Salinas Group, shown in the photographs in Figure 6-8, includes a body of rhyolite, possibly a cryptodome since probable properties were seen at the contacts.

 

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The highest stratigraphic part of the Salinas Group, at least 60 m thick and above the sinters, is cut in the graben in the hanging wall of the East Fault. It comprises lacustrine siltstones and volcaniclastic sandstones. The rocks are plant-rich and contain rare fish fossils and brine shrimp/ostracods (e.g., drill hole CB332).

 

 

 

Figure 6-8: Example Drill log from the Salinas Group

 

Source: Pratt and Gordon, 2019.

 

The Salinas Group includes common mass flow or hydrothermal breccias. Their geometry is frequently unclear; it is uncertain if they are dykes or aprons of phreatic (explosion) breccia ejected from hot springs. Some contain sinter clasts, confirming phreatic eruptions. Underground, the South Ramp is dominated by hydrothermal breccias (Hbx), with polymict clasts up to 0.5 m in diameter. This may be the north margin of a south-dipping diatreme. Successive cross-sections show it extending progressively deeper towards the south.

 

Quaternary basalts (bi), with a felted, trachytic texture, crop out in the north of the Era Dorada property and occur in the low graben on either side of the horst. They are clearly lava flows. Around the village of Cerro Blanco, they in-fill the paleo-topography formed by a large dacite flow dome. It is unclear if this topography is erosional or the original hummocky shape of the dacite flow. The basalts include flow-foliated and autobrecciated types.

 

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The youngest rocks comprise alluvium, and in a few places, modern travertine and tufa occur at springs around the flanks of Cerro Blanco hill (Figure 6-9). The tufa cements colluvial blocks of the siliceous sinter (Salinas Group) are modern and should not be confused with sinter. They imply probable karst formation and dissolution of limestone.

 

 

 

Figure 6-9: Recent travertine exposure

 

Source: Pratt and Gordon, 2019.

 

Discordant igneous intrusions are rare at Cerro Blanco, but a few thin rhyolites (Rp) and aphanitic andesite (ad) dykes are observed.

 

6.3.2Structure

 

The gold mineralization at the Project is hosted within a broadly north-south-striking graben. The East Fault (Figure 6-10), also referred to as the “East Horst Fault” in previous studies, is cut by several drill holes and

 

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observed in the drill core as a broad zone of post-mineral cataclasite developed in Mita siltstone; however, the structure appears to control a linear rhyolite body, suggesting that it was also active during the mineralizing event. This fault may be listric and made up of several strands. Holes CB332 and CB329, in the section below, show narrow wedges of ‘exotic’ lithologies along the fault zone, including limestone (Mls) and conglomerate (Mss). The apparent displacement, shown by the offset of the sinter (Ss), is about 300 m.

 

The immediate footwall of the East Fault, which hosts the gold-bearing quartz veins, is structurally complex. A deep geothermal drill hole (MG-07) shows gold mineralization in the probable down-dip extension of the East Fault at 634-640 m downhole depth.

 

 

 

Figure 6-10: Simplified west-west cross-section across Era Dorada

 

Note: Many drill holes and some lithostratigraphic units and faults were omitted to conserve clarity. Source: Pratt and Gordon, 2019.

 

The Cerro Blanco property has a complex history of faulting. The structural control on mineralization is unusual for low-sulfidation epithermal vein deposits, which normally comprise a single, relatively continuous vein. At Era Dorada, there are sheeted vein swarms that resemble a duplex. Figure 6-11 indicates the typical complexity in an east-west section. Note that the thickness of the Mcv west of the Mat Fault and the Mvo to the east is an artifact of Leapfrog software and is overstated. Veins are shown in red, and faults in white.

 

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Figure 6-11: East-west cross-section of the South zone, Era Dorada looking North

 

Source: Bluestone, 2021.

 

Simplistically, the structural history is comprised of the following:

 

1.Sedimentation of the Mita Group in a basinal to shelf environment, with periodic incursions of calc-alkaline volcanism (mostly waterlain andesitic tuffs and andesite flows, and their volcaniclastic equivalents). Some of the beds appear turbiditic (silt), implying moderate water depth. Some metamorphic clasts imply a metamorphic hinterland.

 

2.A compressive episode formed a series of broadly north-south-striking, west-verging folds cored by Mita Group rocks, in particular, the Mvo and Mcv. These folds were associated with west-verging reverse faults and resulted in local overturned limbs. There may have been a component of strike-slip, with the development of a positive flower structure at the restraining bend in a major north-south strike-slip fault. There is evidence that most of the gold-bearing veins developed at this stage. The controlling structures for the vein swarms are in the footwall of the East Fault and apparently steeper (e.g., the Main Fault, see Figure 6-10).

 

3.Major extensional faulting with downthrows to the east of up to several hundred meters. These include the Ramp and East faults (see above). These faults may have been active during deposition of the Salinas Group (Svc), possibly growth faults. Metamorphic clasts in the Salinas Group imply continued input from a metamorphic hinterland. The offset of Quaternary basalts implies that the faults may still be active (neotectonic). These faults have the greatest surface expression, reflected in the modern topography by the Cerro Blanco ridge and flanking low-relief alluvial plains.

 

Most of the gold-bearing veins are constrained between the Mat Fault in the west and the East Fault, and evidence suggests that most veins at this stage developed along early pre-mineral faults. The Mat fault is interpreted to be a major early structure and hosts the principal footwall vein (VS-101) in the South Zone for

 

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some of its length. The lack of continuity of major veining up into the Salinas suggests that much of the faulting had ceased by the time of the Salinas deposition, except for the Cross, Ramp, and East Faults.

 

Some of these faults may represent syn-volcanic growth faults typical of near-surface epithermal settings that represent shallow, low displacements that manifest as larger pre-mineral faults at depth with increased displacements.

 

In the southeast portion of the South Zone, narrow sub-vertical gold-bearing veins extend into the Salinas and possibly represent a progression from the early compressive to more extensional conditions by the end of the Salinas deposition. Drilling demonstrates that a large chunk of stratigraphy is missing in the area separating the north and south zones of the deposit. This comprises the Mls + Mbt (lower) and Mat. A northwest-striking, southwest-dipping fault (“Upper Mbt Fault”) is inferred. It is unclear if this terminates into the major Ramp Fault or vice versa. The throw on the Upper Mbt Fault seems to decline towards the north, and the stratigraphy is increasingly preserved in the footwall. Together, the Ramp and Upper Mbt faults define a triangular-shaped block that seems to have slid out southwards. Explaining the geometry, in terms of tectonic regime, is difficult, but a reactivated, extensional flower structure is one possible explanation.

 

Faults are difficult to map underground and in drill core because they are largely quite narrow (centimeter scale) and ‘sealed’ by silica; they generally do not form the zones of poor rock quality that typify post-mineral faults (though there are exceptions, for example along the East and Cross faults). This is reflected underground by the general lack of wall rock support. Figure 6-12 shows structural measurements from the underground workings for faults and veins. However, most understanding of the principal faults comes from 3D modeling, based on offsets of the lithostratigraphy and the marker beds.

 

The underground workings display numerous swarms of quartz veins. There are examples of conjugate veins and veins refracting through different lithologies (competency control). Examples are shown in Figure 6-13.

 

The gold-bearing veins at Era Dorada are focused in the footwall to the west of the steep Main Fault (also referred to as the Main Zone); in particular, they are concentrated in the uplifted blocks and west-verging folds of basement volcanic rock (Mcv and Mvo). The Upper Mbt lithostratigraphic unit seems to have been less favorable for veining, explaining the relative gap in veining between the North and South ramps. Likewise, the veins tend to pinch out in the Salinas Group (though some do make it to the surface and carry low grades).

 

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Figure 6-12: Stereograms (equal area) showing poles & great circles for faults & veins

 

Note: All measured underground. Dots on the great circle plots represent slickensides. Source: Pratt and Gordon, 2019.

 

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Figure 6-13: Photographs with sketches of veins exposed underground

 

Source: Pratt and Gordon, 2019.

 

In section view, the veins clearly form lozenge-like duplexes and sheeted swarms, one in the South Ramp, the other in the North Ramp. Figure 6-14 is a cross-section across the South Ramp. Vein wireframes were generated in Leapfrog using core logging, alpha angles (angle between the core axis and vein) in non-oriented drill core, assay data, and underground mapping. They show a distinct branching and converging of relatively shallow veins into a steeper zone (Main Fault). Most veins are also constrained to the footwall of the Ramp Fault and the hanging wall of the steeper Mat Fault.

 

Sheeted veins and lozenge-shaped duplexes are also obvious in the map (plan) view. Figure 6-15 shows a series of horizontal slices at different elevations. The gap between the South and North resource areas mostly comprises the triangular wedge Upper Mbt stratigraphy between the Upper Mbt and Ramp faults. This seems to have been unfavorable for veining.

 

Underground mapping supports the 3D modeling; it shows a similar steepening and converging of veins into the Main Fault / Zone. Individual veins become thicker and more closely spaced along the Main Fault. The way individual veins swing into and intersect with the Main Fault creates ore shoots that plunge approximately 30° south.

 

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Figure 6-14: Annotated, vertical east-west cross-section across the south ramp (looking North)

 

Source: Bluestone, 2020.

 

 

 

Figure 6-15: Horizontal Slices at different elevations through Era Dorada

 

Note: North is up. Red – veins; blue – faults.

Source: Bluestone, 2020.

 

There are some secondary (conjugate) vein directions, but stereograms for sub-areas (Figure 6-16) show consistent patterns: steeper veins are mostly in the east and shallow veins in the west. A swarm of thick, sub-horizontal veins occurs in the immediate footwall of the Ramp Fault. The cumulative thickness of the veins

 

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exceeds 3 m. The flat veins clearly imply reverse (compressive) movement on the Ramp Fault. Clearly, the major faults played an important role in partitioning vein development.

 

 

 

Figure 6-16: Stereograms for more detailed sub-areas in underground mapping

 

Source: Pratt and Gordon, 2019.

 

The stress regime during vein formation can also be calculated from conjugate veins. The stereogram for all quartz veins measured underground shows the intersection between the two principal vein directions is sub-horizontal). The dominant extension direction seems to have been vertical, which is highly unusual; epithermal veins generally develop during horizontal extension. The predominance of horizontal veins in the west supports the idea of vertical extension.

 

Field observations, 3D modeling, and stereograms, therefore, imply that the veins developed during compression rather than extension, at least in the initial stages of mineralization. This fits with the overall compressional geometry of the west-verging folds and reverse faults, later reactivated as normal, extensional faults. Recently discovered steeply dipping/vertical veins in the hanging wall of the south zone possibly record this change from a more compressional to extensional regime during the latter part of the mineralizing event. As some steep veins cut the Salinas Group and the sinters are contemporaneous with hydrothermal activity;

 

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this suggests that the hydrothermal/geothermal activity spanned the change from compressional to extensional tectonics.

 

6.4Deposit Geology

 

6.5Deposit Type

 

The low sulfide content and near absence of base metals in the Era Dorada veins confirm it as a classic hot springs-related, low-sulfidation epithermal deposit. In common with most low-sulfidation deposits, it appears to be linked to compositionally bimodal, basalt-rhyolite volcanism, the hallmark of intra- and back-arc rift settings worldwide. The hydrothermal system seems likely to have been initiated during rhyolite dyke and cryptodome emplacement, at the base of the Salinas unit, with the rhyolitic magma and magmatic input to the mineralizing fluid both being derived from the same deep parental magma chamber.

 

Arc-related low-sulfidation gold deposits occur at the highest crustal levels, most removed from inferred intrusion source rocks. Figure 8-1 shows the generalized deposit model.

 

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Figure 6-17: Generalized deposit model schematic

 

Source: Corbett and Leach, 1998.

 

Adularia-sericite epithermal gold-silver deposits characteristically occur as banded fissure veins and local vein/breccias, which comprise predominantly colloform banded quartz, adularia, quartz pseudomorphing carbonate, and dark sulphidic material termed ginguro bands. Examples of adularia-sericite epithermal gold-silver deposits include Waihi and Golden Cross, Pajingo, Vera Nancy, Cracow, Hishikari, Sado, Konamai, Tolukuma, Toka Tindung, Lampung, Chatree, Cerro Vanguardia, Esquel, El Peñon.

 

At near surficial levels, many are capped by eruption breccias and sinter deposits. Eruption (phreatic) breccias, which form by the rapid expansion of depressurized geothermal fluids, are characterized by intensely silicified matrix and generally angular fragments, including sinter, host rock, and local surficial plant material. Although sinter deposits formed distal to fluid upflows commonly associated with eruption breccias, sinters tend to be barren with respect to gold but may be anomalous in other elements such as boron, arsenic, and antimony.

 

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Although cooling and traditional boiling models still hold for the deposition of gangue minerals (adularia, quartz pseudomorphing platy calcite, and chalcedony) and some gold, mixing of rising pregnant fluids with oxygenated or collapsing acid sulfate (low pH), groundwater is also favored as a mechanism for the development of characteristic bonanza gold-silver grades. Adularia-sericite vein systems are silver-rich, with gold-to-silver ratios greater than 1:10 being common.

 

Wall rock alteration formed as halos to veins occurs as sericite (illite) grading to peripheral smectite clays with associated pyrite and chlorite, and this alteration grades to more marginal chlorite-carbonate (propylitic) alteration. Low-temperature acid waters developed by the condensation of volatiles in the vadose zone contribute towards the formation of surficial acid sulfate alteration comprising silica (chalcedony, opal), kaolin, and local alunite, and these acid sulfate waters are interpreted to collapse to deeper levels and so aid in mineral deposition.

 

Structure and host rock competency are important mineralization controls in adularia-sericite vein systems. High-grade mineralized shoots often develop in dilational jogs or flexures in through-going veins where veins of greater thickness and higher gold grade develop and the intersections of fault splays. Bonanza-grade material may also develop at preferred sites of fluid quenching at rock competency changes. Recent studies (e.g., Rhys et al., 2020) attest that fault systems in very shallow epithermal systems characterized by sinter, lacustrine sediments, and hydrothermal breccias, similar to Era Dorada may represent syn- volcanic low-displacement growth faults that manifest as larger displacement pre-mineral faults at depth.

 

The connection between modern hot spring deposits and ancient hydrothermal systems, some with gold mineralization, has long been recognized (Lindgren, 1933). Epithermal mineral deposits are defined as those that develop close to the Earth’s surface (within 1,000 m). They developed from fluids like those in modern geothermal systems. Sillitoe and Hedenquist (2003) defined the three types of epithermal deposits: high, intermediate, and low sulfidation. The low-sulfidation variant commonly occurs in rift settings, with bimodal volcanism in young, often Tertiary, volcanic arcs (e.g., Henley and Ellis, 1983). It is commonly associated with maar volcanoes, diatremes, and felsic flow domes.

 

Era Dorada shows all the characteristics of a completely preserved, non-eroded epithermal deposit. The occurrence of hot springs (sinters, silicified reeds, pisoliths) directly above the presumed feeder veins at Era Dorada implies a high water table and swampy conditions (cf. McLaughlin, California). In areas of high topographic relief, outflow springs (sinter) are usually found several kilometers from the upflow zones. The widespread occurrence of lacustrine and fluvial clastic sediments in the Salinas Group and accretionary lapilli, typical of water-rich pyroclastic surges, supports this interpretation. Sedimentation probably kept up with subsidence. Mudstone dykes and geopetal structures—open fractures filled by horizontally bedded chalcedonic and Sulfide-rich sediment—reinforce the interpretation.

 

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6.6Era Dorada Deposit Geology

 

The Era Dorada deposit is a classic hot springs-related, low-sulfidation quartz-adularia-calcite vein system. It is localized along a complex fault intersection created during the late Miocene-Pliocene tectonic extension within the active Central American volcanic arc. Local igneous activities that drove the Era Dorada hydrothermal system include a vesicular andesite dike swarm and mineralization stage rhyolite/dacite flow dome eruption and cryptodome intrusion.

 

The Era Dorada vein systems are best developed (widest and most continuous) between the 300 masl to 500 masl elevation ranges. Principal host rocks include a lithic tuff—calcareous shallow marine-volcaniclastic sequence and, to a lesser extent, the overlying volcaniclastic-hydrothermal breccia sequence of probable Pliocene age. Vein zones often appear to transition to barren calcite beneath the ±300 m elevation in the northern half of the deposit. To the south, high-grade quartz-adularia-calcite vein zones continue at least another 100 m down to 200 m elevation. Some veins remain open at depth.

 

Massive chalcedonic silicification, referred to as a “silica cap,” dominates the conglomerates of the Salinas unit. Silica-flooded volcaniclastics and phreatic breccia are interbedded with chalcedonic silica sinter from the present surface to depths of ±100 m. Silicification also occurs in the underlying Mita as irregular envelopes, up to several meters wide, around the main veins as well as in the upper part of the limestone horizon as jasperoid. The red-bed siltstone is partially bleached and altered to a grey-green, illite, and smectite-bearing rock. Chlorite, in addition to illite and smectite, is a prominent alteration mineral in the ignimbrite, where it is concentrated in the fiamme.

 

Wall rock alteration, to a large extent, determines geotechnical rock hardness and presents contrasting resistivity and electrical chargeability characteristics that could be exploited across the district in the search for new gold occurrences beneath thin colluvial or basalt cover.

 

6.7Mineralization

 

The Era Dorada gold deposit occurs within a large hydrothermal alteration zone covering an area of about 5 km long and 1 km wide. This zone exhibits the effects of strong, pervasive hot spring-type hydrothermal alteration.

 

Gold mineralization is hosted within a broadly north-south striking sequence of westerly-dipping siltstones, sandstones, and limestones (Mita Group) that are capped by silicified conglomerates and argillaceous sediments with contemporaneous dacite/rhyolite flow domes or cryptodomes (Salinas Unit). The Salinas rocks are syn-mineral and believed to have accumulated progressively in a low-relief graben characterized by a shallow groundwater table. The Salinas conglomerate was presumably derived by erosion of the flanking horst blocks as relief was created during active faulting. The topographic inversion required

 

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to explain the current prominent position of the graben fill is ascribed to the silicic character of the Salinas unit and its consequent resistance to erosion.

 

The west and east sides of the Era Dorada ridge consist of flat agricultural plains characterized by Quaternary basalts, interbedded with boulder beds and sands. These rocks also appear down-faulted to lower elevations, implying major post-mineral extensional movements on such faults, and they may be neotectonic (active).

 

The current gold resource occurs under a small hill and is confined within an area of about 400 m x 800 m. Gold and silver occur almost exclusively in quartz-dominated veins of low-sulfidation epithermal origin and in low-grade disseminated mineralization within the Salinas conglomerates and rhyolites. The highest grades are hosted by high to low-angle banded chalcedony veins, locally with calcite replacement textures.

 

Gold-bearing structures in the Project area extend 3 km to the northwest of the gold deposit and occur largely confined within the hydrothermal alteration zone. Exposures are poor and locally covered by alluvium and post-mineral rocks. Gold-bearing structures extend at least 1 km south and southwest of the deposit under valley fill and post-mineral rocks.

 

Geothermal well MG7, located about 0.5 km east of the deposit, encountered a 27 m zone averaging 6.3 g Au/t and 22 g Ag/t at a depth of 634 m. The upper 6 m of this zone averages 23.9 g Au/t and 79 g Ag/t. Although the geometry is uncertain and the sampling methodology of the drill cuttings cannot be determined, possibly this vein material was caught up in a fault crush zone/splay within the East Fault (much like the other exotic lithologies seen within the fault zone), or conversely, represents a separate mineralized system distinct from the main deposit.

 

6.7.1Vein Zones

 

Petrographic descriptions of four vein zones by Economic Geology Consulting (Thompson et al., 2006) concluded that the veins consist of crustiform banded chalcedony, quartz, adularia, calcite, sulfides, and visible gold. The samples represent a range of almost 300 m in elevation. Bladed calcite or pseudomorphs after bladed calcite (lattice blade texture) were observed in all four samples. Bladed calcite is a rapid depositional texture, common when calcite precipitates from boiling fluids. A wide variety of recrystallization textures in quartz and chalcedony may also indicate changing fluid conditions and periodic boiling. Figure 6-18 shows a high-grade intercept in drill hole CB-20-430 with banded chalcedony-adularia-acanthite and visible gold that assayed 144 g Au/t and 282 g Ag/t.

 

Observations suggest that mineralization occurred as one principal multi-stage event as banded vein material, dominated by cryptocrystalline and originally amorphous silica phases (jigsaw quartz and chalcedony) characteristic of both the north and south zone vein swarms. Colloform banding with gel-like precursor textures is common, and observations from drill core suggest that banding is characteristic of high-grade zones, with coarser crustiform and crystalline bands more associated with lower-grade veins. Higher grades are associated

 

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with fine-grained (<100 µm) electrum, kustelite, and acanthite concentrated in bands of fine- to very fine-grained jigsaw quartz (crystallized amorphous silica, Albinson, 2019). Gold-silver minerals are accompanied by the rare presence of tetrahedrite and chalcopyrite.

 

Repetitive “crack and seal” pulses and associated boiling/flashing events very close to the paleosurface are suggested as the main mechanisms for precious metal deposition. The higher-grade, often bonanza-grade core intersections with coarser and more abundant sulfides, electrum, and free gold appear to represent an earlier series of events. Multistage banding can be very finely repetitive down to 5 to 10 µm widths for individual bands. Soft sediment-type deformation is commonly visible in the bands with mamillary colloform bands deformed into flame-like textures due to the deformation of the bands by turbulent fluid flow. Sulfides and electrum are present mainly in the fine- or very fine-grained jigsaw quartz bands. Adularia-rich bands are not easily visible with the hand lens and are very fine-grained.

 

 

 

Figure 6-18: High-grade drill hole intercept hole CB20-430 – 144 g/t Au, 282 g/t Ag (227.3 to 228.9 m)

 

Source: Bluestone, 2020.

 

The lack of inter-stage hydrothermal brecciation and coarse-grained primary quartz textures suggest that the mineralizing event was a fairly short-lived event that occurred very close to the paleosurface. The lack of post-mineral structural displacement of veins and distribution of high grades over a +300 m vertical profile attest to the pristine nature of the veins.

 

Underground observations include the following:

 

·Vein zones are best developed throughout the model between elevations of 300 and 500 m. This elevation range roughly coincides with the Mcv contact beneath and the Salinas contact above. Thus, the principal host rocks are the Mita Group sandstones, calcareous sediments, and overlying tuffs.

 

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·The quartz veins at Era Dorada occur mainly within Mita sediments and Mcv tuffaceous rocks. These moderate to steep veins are associated with a subsidiary conjugate set of low-angle veins. The majority of veins appear to stop at the Salinas contact, with the exception of sub-vertical veins in the southeast part of the south zone that cut the Salinas and continue to surface.

 

·Vein zones occur as two upward-flared arrays that appear to converge downwards and merge with basal master veins around the contact with the Mcv. The south zone vein array is the better-formed

 

·High gold grades locally persist at least down to the 200 m elevation, notably in the southern third of the model, where at least one vein merges with the main footwall feeder structure.

 

·In several locations north of 1,587,400N, drill holes pass beneath high-grade quartz veins but encounter only massive barren calcite. This is an indication that the bottoms of productive veins have been found at those locations. Within vein zone envelopes, individual veins do not form a random stockwork but tend to run parallel or sub-parallel to the main structural trends.

 

The definition of economic mineralization depends on the vein thickness, grade, and spacing. The structural control of the veins is discussed above. Most individual veins exposed in the underground workings do not exceed 1 to 2 m; much thicker veins, up to 7 m width, do appear in the vicinity of the north zone ramp (Figure 6-19) and in deeper levels of the south zone. Closely spaced veins or zones of convergence form wide zones of high-grade mineralization (Figure 6-19).

 

 

 

Figure 6-19: View of veins VN-05, 06, 07 in the north ramp underground workings

 

Note: Section assayed 20.4 m grading 18.9 g Au/t and 33.2 g Au/t.

Source: Bluestone, 2020.

 

Figure 6-20 shows vein textures associated with gold mineralization; they include bladed calcite, a classic indicator of boiling fluids, subsequently replaced by quartz or leached to give a skeletal framework.

 

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Other classic textures include crustiform banding, bands of cream-pinkish euhedral adularia, and quartz with minor dark grey silver Sulfides/sulphosalts.

 

Inspection of vein textures suggests that gold and silver were introduced as one major event of multistage finely banded veining (originally amorphous silica) with subordinate bands of platy calcite that is mostly pseudomorphed to cryptocrystalline silica phases.

 

 

 

Figure 6-20: Examples of vein textures from Era Dorada

 

Source: Bluestone, 2020.

 

Many veins and siliceous rocks (rhyolite/dacite) at Era Dorada display siliceous mudstone/sandstone dykes. There are also common geopetal structures, late cavities filled by horizontally banded siliceous sediments of hydrothermal origin mixed with vein gangue (Figure 6-21). These “fossil spirit levels” indicate proximity to the paleosurface and are confirmed by the presence of sinter immediately above.

 

It is unusual to see epithermal veins developed immediately beneath sinter, although other examples do exist (e.g., McLoughlin, California), implying the topography at the time of mineralization was low and the water table was very high. This is supported by the presence of accretionary lapilli in the Salinas Group and Mbt siltstones; they are typical of wet phreatic-dominated eruptions and pyroclastic surges. Diatremes and rhyolite flow domes are also typical in this environment.

 

In summary, the principal control on gold mineralization at Era Dorada was probably the boiling level in a hydrothermal system. The best grades are associated with boiling textures. At many low-sulfidation epithermal

 

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deposits, the vertical interval of economic grade is restricted to the former boiling level. This can be less than 100 m. These boiling levels form flat ore shoots. There are occurrences of high gold grade down to 640 m (downhole depth) in a geothermal hole (MG-07).

 

 

Figure 6-21: Example of geopetal structure

 

Source: Pratt and Gordon, 2019.

 

6.7.2Disseminated Mineralization

 

The Salinas unit shows widespread and low-grade disseminated gold mineralization associated with weak to strongly silicified polymictic conglomerates and altered rhyolite breccias and flows. Mineralization grading of 0.2 to 2 g/t Au is pervasive and present in variably silicified bedded conglomerates and appears to be driven by intrusive rhyolite dykes and breccias (Figure 6-22). Locally, parts of the base of the Salinas are marked by an aphanitic rhyolite body, probably a cryptodome, given it is underlain by narrow rhyolite dykes. The thicker Sinter horizons do not contain significant gold values, nor do strongly argillic-altered lithologies and fault gouge zones.

 

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Figure 6-22: Salinas Unit – examples of disseminated mineralization rock types, Salinas Unit

 

Source: Bluestone, 2020.

 

6.7.3Hydrothermal Alteration

 

Many low-sulfidation epithermal vein deposits have significant, mechanically weak halos of illite/smectite + pyrite + sphene/leucoxene; however, the wall rocks at Era Dorada are generally only weakly clay altered and have a very low Sulfide content (Figure 6-23). Most clay alteration is concentrated along some late faults, for example, the East and Cross faults, and within some of the hydrothermal breccias, particularly the phreatic breccias in the Salinas Group.

 

A study using drill core hyperspectral imaging spectroscopy in the 500 nm to 2,500 nm wavelength range and detailed petrographic, SEM, and EDS studies revealed two paragenetic stages of vein formation (Savinova, 2020). The main auriferous veins consist of multi-stage crustiform and colloform bands that are characterized by paragenetic Stage 1 equilibrium assemblage of quartz (chalcedony)-adularia-calcite- ankerite. Sulfides are located mostly in ginguro bands that consist of fine-grained pyrite, chalcopyrite, tetrahedrite, and acanthite. Stage 2 of the paragenesis is characterized by intense overprinting of the quartz-adularia veins by montmorillonite and interstratified illite. Locally, bladed calcite is replaced by quartz. Hydrothermal alteration in the proximal zone of the sedimentary and volcanoclastic wall rocks is characterized by quartz-adularia-illite-montmorillonite. Wall rock-hosted illite suggests a temperature of formation >230°C. The distal alteration zone is marked by illite-chlorite-calcite.

 

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Figure 6-23: Vertical alteration profile through Era Dorada

 

Source: Savinova, 2020.

 

Silicification is widely developed within the Salinas and more selectively in the underlying Mita Group, where it occurs as irregular envelopes, up to several meters wide, around the main veins as well as in the upper part of the limestone horizon as jasperoid. The most impressive alteration feature at Era Dorada is the large “silica cap” hosted in the Salinas sediments, typically beginning at or below 400 m elevation and continuing upward to the surface. Most silica is directly related to hot spring activity; the sinters and pisolithic beds contain abundant silica (although it is possible that some had carbonate precursors). However, there are also numerous beds of sandstone, conglomerate, and mass flow deposits in the Salinas Group that are highly siliceous and locally flooded by chalcedony and fine-grained pyrite. These rocks are black when fresh, white, and limonite stained when oxidized. Exposures around the Era Dorada ridge show that this silicification can be very capricious and replaced abruptly and laterally by smectite-rich clay alteration.

 

Where the paleo-groundwater table was several meters below the paleosurface, a conglomerate in its immediate vicinity was silicified, and the vadose zone above it was subjected to steam-heated alteration. The steam-heated alteration, containing cristobalite, kaolinite, and alunite (an advanced argillic assemblage), was the product of acidic solutions formed by the condensation of ascendant H2S-bearing steam into downward-percolating groundwater. The overall result is an interlayered sequence of sinter, silicified conglomerate, and steam-heated alteration.

 

Many faults at Era Dorada are sealed by silica and are pre-mineral. Examples are shown in Figure 6-24.

 

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Figure 6-24: Examples Of Sealed, Silicified Fault Zones

 

Source: Pratt and Gordon, 2019.

 

The boiling hydrothermal fluids that formed the Era Dorada vein system produced an even larger volume of intensely altered wall rock. Alteration types and zoning are typical of low-sulfidation epithermal systems. The remnant sinter above the deposit suggests that the Era Dorada system remains largely intact.

 

Silicification continues locally down to 300 m elevation along fault zones and in favorable rock types. Overall, the Era Dorada silica cap averages 400 m wide and is up to 150 m deep for at least a kilometer in strike. Within 50 m to 100 m of the surface, silicification is manifested by opaline silica flooding in the fragmental Svc and Rp units. At depth, very fine-grained quartz replacement of Mita Group calcareous sediments (locally forming jasperoid) and tuffs dominate. The Mcv crystal lithic tuff is generally only silicified near contacts with overlying sediments and along fault zones.

 

Silicification typically yields outward to moderate to strong sericitic alteration above 400 or 450 m elevation. At deeper levels, silicified zones grade outward and downward into large volumes of clay-sericite- pyrite±calcite alteration in Mita Group sediments and tuffs. Pyrite contents are commonly in the range of 1-3%, locally reaching 5%.

 

The Mcv is pervasively sericite-chlorite-pyrite±calcite altered virtually everywhere it has been drilled. Sericite dominates closer to mineralized faults and higher. Chlorite-calcite dominates outward and at depth. Pyrite is ubiquitous but generally less than 0.5%.

 

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7EXPLORATION

 

As of the end of 2021, Bluestone had drilled approximately 267 holes for a total of 45,725 m on the Era Dorada property since acquiring it from Goldcorp. Table 7-1Table 7-1 summarizes historical drilling on the property.

 

Table 7-1: Drilling summary

 

Year Company Holes Drilled Meters
1998 Mar-West 9 1,340
1999 Glamis 48 7,074
2000 Glamis 18 3,525
2002 Glamis 23 6,525
2004 Glamis 42 9,370
2005 Glamis 120 29,065
2006 Glamis 67 15,129
2007 Goldcorp 47 12,373
2008 Goldcorp 2 586
2009 Goldcorp 1 140
2010 Goldcorp 10 2,277
2011 Goldcorp 28 5,898
2012 Goldcorp 96 21,370
2017 Bluestone 8 2,324
2018 Bluestone 74 13,993
2019 Bluestone 61 8,403
2020 Bluestone 74 15,172
2021 Bluestone 50 5,833
Total 778 160,397

 

Source: Kirkham, 2021.

 

Figure 7-1 shows a plan view of drill hole locations. Figure 7-2 and Figure 7-3 show representative section views of the drilling along with gold assay data and topography.

 

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Figure 7-1: Plan view of drill hole locations

 

Source: Kirkham, 2021.

 

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Figure 7-2: Section View A-Aʹ (Azimuth 110°)

 

Source: Kirkham, 2021.

 

 

 

Figure 7-3: Section view B-Bʹ (azimuth 110°)

 

Source: Kirkham, 2021.

 

7.1Goldcorp & Glamis Drilling (Pre-2017)

 

Prior to Bluestone’s ownership, reverse-circulation (RC) and diamond drilling (DD) was carried out. Many early holes were collared using RC size core before switching to NQ size core. Collar data from these historical programs was surveyed with a differential global positioning system (GPS), and down-hole survey measurements were taken with either a single-shot Sperry-Sun camera system or a multi-shot Flexit instrument.

 

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Many of the earlier drill holes by previous operators were not drilled perpendicular to the strike and dip of the veining, and therefore, drilled widths of many veins were not representative. The most common vein intersections occur from between 0° and 60° to the core access. These intervals are thought to belong to steep to moderately dipping vein sets. These core intervals would be longer than the true thickness of the actual veining. Intersections ranging from 60° to 90° to the core axis are less common and are believed to belong to flat to near-flat vein structures. These vein intervals would be closer to the true thickness of the veining but still longer than the true thickness. Only vein intervals drilled perpendicular to the strike and dip of the veining would represent the true thickness of the vein. Based on previous reports from Glamis Gold, the ratio to the true thickness of the vein on average is about 1.73 (i.e., every 1.7 m represents 1 m of true vein thickness).

 

7.2Data Validation

 

Historical core logging, sampling, and quality assurance/quality control (QA/QC) procedures were first reviewed and documented by Golder in 2014. Ten core samples were collected from one-quarter sawn NQ core, and selected drill hole collars were surveyed using a GPS. Assayed gold and silver grades were found to be consistent with those reported by Goldcorp. Golder was satisfied that the drill hole data was collected in a manner consistent with industry best practice standards.

 

As part of the core logging data verification, Golder compared a selection of core logs against half-core stored at the project site. Five half-core drill holes were reviewed from the North and South deposits. The Microsoft Excel files were reviewed first, and drill holes were selected that represented the typical mineralization style for each deposit. In addition, 10 verification samples were taken from these drill holes. Each verification sample was a half-core sample sawed into quarters, with one-quarter sample sent for analysis and the other returned to the core racks. Table 7-2Table 7-2 on the following page summarizes the samples selected for core logging review and verification sampling.

 

Samples were sawed and bagged under Golder's supervision and were transported off-site via helicopter and plane to Canada and then by ground transportation to ALS Chemex Laboratories in Sudbury for sample preparation and analysis. A comparison of the Excel files against the drill core indicated an excellent match between the core logs and the retained core. Table 7-3Table 7-3 provides a list of the drill hole collar surveys completed by Golder.

 

Eight drill sites were visited, with multiple drill holes located at some sites. Casings had been removed for most drill holes. The data collected was a mixture of pre-Goldcorp drill holes (2006 or earlier) and drilling completed by Goldcorp during 2010 and 2011. All drill holes from the surface were grouted to prevent water flow into the underground workings.

 

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Table 7-2: Verification samples

 

Drill Hole ID Duplicate Sample No. Original Sample No From (m) To (m) Deposit Metal Analysed Rock Type
CB-152 205873 82225 128 129 North Au, Ag Lapilli Tuff
CB-152 205874 82226 129 130 North Au, Ag Lapilli Tuff
CB-200 205884 407101 156 157 South Au, Ag Quartz Tuff
CB-200 205885 407102 157 158 South Au, Ag Quartz Tuff
CB-241 205891 404849 111.4 112.6 South Au, Ag Conglomerate
CB-241 205892 404850 112.6 113.5 South Au, Ag Fault
CB-254 205895 414397 100.5 102 South Au, Ag Volcaniclastic Sediments
CB-254 205896 414398 102 103.5 South Au, Ag Volcaniclastic Sediments
CB-10-15 205871 435941 135 136.23 North Au, Ag Lapilli Tuff
CB-10-15 205872 435943 136.23 137.46 North Au, Ag Lapilli Tuff

 

Source: Goldcorp, 2014.

 

Table 7-3: Drill hole collar survey (NAD 27 Zone 16N)

 

Drill Hole ID Golder Cerro Blanco
Easting Northing Easting Northing
C 10 08 212015.1 1587867 212009 1587748
C 11 12 211906.8 1587714 211904 1587605
C 11 15 211969.7 1587769 211966 1587655
C 11 18 211866.4 1587405 211873.2 1587297
C 11 21 211901.6 1587414 211898.9 1587307
C 151 212025.1 1587821 212020.8 1587707
C 247 211985.5 1587315 211978.8 1587202

 

Source: Goldcorp, 2014.

 

Approximately 5% of the drill holes (20 holes) were subjected to data verification checks by Golder. The 20 selected holes, summarized in Table 7-4, included a variety of historical data as well as some of the more recent holes. The data verification checks consisted of the following:

 

·comparison of final assays to the original laboratory certificates

 

·analysis of external laboratory duplicate assays by generating XY scatterplots

 

·review of downhole survey measurements to identify anomalous changes to hole orientation.

 

For the 20 holes reviewed, the comparison of final assays to the original assay certificates did not identify any material differences in assay values.

 

External laboratory duplicate assays were reviewed to assess the reliability of the primary assay laboratory. XY scatterplots were generated for each of the 20 holes. With the exception of a few outliers, the majority of the data compared well. Figure 7-4 illustrates an example of the XY scatterplots used to compare assay results.

 

Table 7-4: Drill holes selected for data verification

 

Drill Hole IDs
CB-012 CB-200
CB-016 CB-227
CB-063 CB-244
CB-078 CB-247
CB-095 CB-305
CB-10-02 CB-309
CB-120 CB-314
CB-142 CB-345
CB-146 CB-357

 

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Drill Hole IDs
CB-151 CB-362

 

Source: Goldcorp, 2014.

 

 

 

Figure 7-4: Example of XY scatterplot for hole CB34

 

Source: Goldcorp, 2014.

 

7.3Bluestone Drilling (2017-2021)

 

Drilling completed by Bluestone between 2017 and 2020 was a combination of surface and underground diamond core drilling. Underground channel sampling was also performed and included in the resource estimation.

 

Drills were operated by Continental Drilling of Guatemala. The surface drilling was performed using two Hydracore 1000 portable drill rigs, one of which was replaced later in the program by a Boart Longyear LM-75 belonging to Bluestone, which was later converted for underground drilling. During the height of the drill program, five LM-75s were operative. Drill holes were developed by drilling a larger diameter (HQ) core at the early stage of the hole and then decreasing to NQ and/or BQ size if the drilling conditions became difficult.

 

Core recoveries were high, and by utilizing several drill core sizes, Bluestone was able to ensure drill hole target completion. To date, 89 holes have been drilled from the surface and 128 holes from underground.

 

Drill hole collars were surveyed using a total station (coordinate system UTM NAD 27 Zone 16N). In-hole drill surveying for azimuth and dip was completed using the Reflex EZ-Shot system approximately every 25 m down-hole. Orientation of the drill core was performed throughout Bluestone’s drill program using Reflex ACT III downhole survey equipment.

 

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7.4Significant Assay Results

 

Table 7-5 provides a selection of significant drill hole intervals from the Era Dorada drill hole database. Drill hole intervals are reported as actual core lengths, and many may not represent the true thickness.

 

Table 7-5: Gold & silver samples from the drill hole database

 

Hole Company From To Length (m) Au (g/t) Ag (g/t)
CB-012

Mar-West/

Glamis

99.50 108.50 9.00 13.7 46.5
CB-012

Mar-West/

Glamis

141.50 147.50 6.00 12.9 75.8
CB-012

Mar-West/

Glamis

195.50 198.50 3.00 3.0 8.0
CB-012

Mar-West/

Glamis

236.00 237.50 1.50 13.0 6.0
CB-016

Mar-West/

Glamis

192.55 195.35 2.80 3.3 0
CB-063 Glamis 88.50 99.00 10.50 4.4 25.7
CB-063 Glamis 114.00 126.00 12.00 3.2 21.0
CB-063 Glamis 183.00 186.00 3.00 7.7 20.0
CB-063 Glamis 196.50 199.50 3.00 4.2 25.0
CB-063 Glamis 207.00 210.00 3.00 18.7 20.0
CB-063 Glamis 225.00 228.00 3.00 37.3 75.0
CB-063 Glamis 241.50 244.50 3.00 5.1 3.5
CB-078 Glamis 158.20 161.40 3.20 3.4 4.1
CB-078 Glamis 242.10 245.10 3.00 3.5 4.7
CB-078 Glamis 248.10 273.75 25.65 66.1 42.2
CB-078 Glamis 299.25 303.75 4.50 4.7 17.7
CB-078 Glamis 338.25 345.75 7.50 10.8 17.4
CB-095 Glamis 155.00 158.00 3.00 3.7 204.9
CB-095 Glamis 179.00 182.00 3.00 17.8 7.4
CB-095 Glamis 233.00 236.00 3.00 88.0 98.6
CB-10-02 Goldcorp 117.50 120.30 2.80 14.7 79.5
CB-10-02 Goldcorp 135.75 139.50 3.75 12.9 91.8
CB-10-02 Goldcorp 146.00 149.00 3.00 9.5 79.6
CB-10-02 Goldcorp 168.86 173.00 4.14 26.2 144.8
CB-10-02 Goldcorp 197.00 200.00 3.00 20.3 19.9
CB-120 Glamis 219.00 238.50 19.50 17.5 20.3
Hole Company From To Length (m) Au (g/t) Ag (g/t)
CB-120 Glamis 246.00 249.00 3.00 8.8 20.6
CB-142 Glamis 163.50 171.50 8.00 16.0 72.2
CB-142 Glamis 196.20 204.50 8.30 19.2 11.7
CB-142 Glamis 302.75 306.00 3.25 19.3 14.3
CB-146 Glamis 80.30 86.00 5.70 14.0 196.8
CB-146 Glamis 109.00 112.40 3.40 10.3 78.9
CB-146 Glamis 118.90 130.00 11.10 70.4 226.3
CB-146 Glamis 139.00 143.00 4.00 12.4 35.4
CB-146 Glamis 149.00 152.00 3.00 3.7 8.0
CB-146 Glamis 156.00 159.00 3.00 21.1 30.6
CB-146 Glamis 182.00 185.00 3.00 4.2 2.5
CB-151 Glamis 162.40 165.50 3.10 25.6 152.8
CB-151 Glamis 172.90 179.30 6.40 13.6 24.7
CB-151 Glamis 327.50 330.50 3.00 5.0 5.5
CB-200 Glamis 117.00 120.00 3.00 5.7 26.0
CB-200 Glamis 144.00 147.00 3.00 5.0 13.0
CB-200 Glamis 152.00 161.00 9.00 7.5 13.6

 

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Hole Company From To Length (m) Au (g/t) Ag (g/t)
CB-200 Glamis 165.00 168.50 3.50 16.7 212.9
CB-227 Glamis 117.34 124.96 7.62 15.4 20.6
CB-227 Glamis 131.00 134.00 3.00 5.6 22.0
CB-244 Glamis 90.00 99.00 9.00 10.3 57.0
CB-244 Glamis 139.50 142.50 3.00 4.2 4.0
CB-244 Glamis 234.00 237.00 3.00 22.5 21.0
CB-247 Glamis 135.00 138.00 3.00 3.5 25.5
CB-247 Glamis 159.00 162.00 3.00 4.0 4.5
CB-247 Glamis 231.00 234.00 3.00 6.8 15.7
CB-247 Glamis 240.00 243.00 3.00 28.6 98.5
CB-305 Glamis 86.00 90.00 4.00 5.0 9.5
CB-305 Glamis 138.00 141.50 3.50 5.5 21.3
CB-309 Glamis 128.50 132.00 3.50 3.5 8.6
CB-309 Glamis 183.00 186.70 3.70 130.1 304.6
CB-309 Glamis 193.50 196.50 3.00 40.3 17.0
CB-314 Glamis 99.50 102.50 3.00 5.3 11.0
CB-314 Glamis 111.50 119.50 8.00 8.3 19.9
CB-314 Glamis 124.50 127.50 3.00 24.2 113.6
CB-314 Glamis 131.50 134.50 3.00 13.6 30.7
CB-314 Glamis 140.50 143.50 3.00 11.8 45.0
CB-314 Glamis 151.50 154.50 3.00 3.7 15.0
CB-314 Glamis 175.50 178.50 3.00 85.6 386.9
CB-314 Glamis 186.00 189.00 3.00 4.2 12.5
CB-345 Glamis 231.70 234.70 3.00 13.1 20.8
CB-345 Glamis 315.50 318.50 3.00 5.8 6.7
CB-357 Glamis 63.00 66.00 3.00 5.5 33.3
CB-357 Glamis 140.00 143.00 3.00 3.4 2.7
CB-357 Glamis 159.00 162.50 3.50 4.0 2.7
CB-357 Glamis 184.00 187.00 3.00 3.6 22.0
CB-357 Glamis 192.50 195.50 3.00 46.4 126.3
Hole Company From To Length (m) Au (g/t) Ag (g/t)
CB-357 Glamis 200.00 206.20 6.20 12.6 6.3
CB-357 Glamis 217.50 220.80 3.30 4.3 5.0
CB-362 Glamis 128.50 131.50 3.00 4.2 6.0
CB-362 Glamis 219.00 222.20 3.20 4.5 6.0
CB17-376 Bluestone 221.90 224.40 2.50 17.1 33.0
CB18-386 Bluestone 243.80 246.47 2.63 5.1 5.6
CB18-388 Bluestone 37.70 41.00 3.30 8.6 3.5
CB18-389 Bluestone 104.70 110.00 5.30 7.9 35.1
CB18-390 Bluestone 164.27 169.57 5.30 16.0 29.1
CB18-393 Bluestone 253.60 261.50 7.90 16.5 18.4
CB18-394 Bluestone 110.60 128.00 17.40 7.0 65.2
CB18-395 Bluestone 46.30 51.00 4.70 5.8 4.2
CB18-396 Bluestone 103.08 108.15 5.07 7.1 24.7
CB18-396 Bluestone 167.14 181.41 14.27 16.2 20.6
UGCB18-71 Bluestone 0.00 27.69 27.69 5.5 17.1
UGCB18-71 Bluestone 0.00 27.69 27.69 5.5 17.1
UGCB18-72 Bluestone 88.10 90.00 1.87 7.6 23.5
UGCB18-73 Bluestone 6.00 23.00 17.00 5.1 17.2
UGCB18-73 Bluestone 37.19 43.13 5.94 5.2 10.3
UGCB18-73 Bluestone 13.20 16.85 3.65 19.3 59.4
UGCB18-74 Bluestone 37.62 41.23 3.61 9.0 28.5
UGCB18-74 Bluestone 54.40 56.39 1.99 21.3 63.4
UGCB18-75 Bluestone 45.72 51.22 5.50 7.3 60.9
UGCB18-76 Bluestone 12.61 47.10 34.49 5.8 18.6
UGCB18-76 Bluestone 12.61 16.53 3.92 26.8 84.4
UGCB18-79 Bluestone 11.31 20.82 9.51 5.6 33.9

 

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Hole Company From To Length (m) Au (g/t) Ag (g/t)
UGCB18-80 Bluestone 47.77 53.25 5.48 9.3 105.3
UGCB18-80 Bluestone 85.95 88.47 2.52 13.9 85.2
UGCB18-81 Bluestone 100.50 105.07 4.57 20.8 46.9
UGCB18-81 Bluestone 122.18 125.20 3.02 11.2 13.1
UGCB18-82 Bluestone 71.16 81.18 10.02 15.0 32.5
UGCB18-84 Bluestone 53.33 56.08 2.75 44.7 39.9
UGCB18-85 Bluestone 52.34 59.12 6.78 24.6 92.8
UGCB18-85 Bluestone 70.05 71.13 1.08 21.2 60.9
UGCB18-86 Bluestone 23.50 30.50 7.00 17.2 94.9
UGCB18-86 Bluestone 33.35 37.19 3.84 9.1 28.9
UGCB18-86 Bluestone 43.55 51.81 8.26 32.7 79.6
UGCB18-87 Bluestone 97.74 98.81 1.07 16.0 26.8
UGCB18-88 Bluestone 43.00 52.20 9.22 9.8 29.9
UGCB18-88 Bluestone 62.20 64.20 2.00 9.8 35.7
UGCB18-89 Bluestone 50.72 65.72 15.00 16.7 105.4
UGCB18-89 Bluestone 92.01 101.37 9.36 14.3 68.5
UGCB18-91 Bluestone 12.90 15.85 2.95 17.9 27.6
UGCB18-92 Bluestone 36.80 58.20 21.40 9.6 34.9
UGCB18-92 Bluestone 112.30 117.60 5.40 12.8 10.8
UGCB18-93 Bluestone 10.30 11.30 1.00 24.5 32.2
UGCB18-94 Bluestone 98.10 100.30 2.20 7.2 15.7
Hole Company From To Length (m) Au (g/t) Ag (g/t)
UGCB18-95 Bluestone 6.40 7.60 1.20 8.9 49.2
UGCB18-95 Bluestone 14.10 15.60 1.50 12.2 27.3
UGCB18-96 Bluestone 39.40 52.40 13.00 11.5 48.6
UGCB18-96 Bluestone 56.40 61.40 5.00 7.1 30.5
UGCB18-98 Bluestone 108.20 110.60 2.30 9.9 8.7
UGCB18-98 Bluestone 115.20 116.20 1.00 28.6 112.0
UGCB19-126 Bluestone 32.20 43.00 10.20 13.1 25.0
UGCB19-143 Bluestone 57.00 66.00 9.00 8.4 53.2
UGCB19-144 Bluestone 98.80 106.70 7.50 19.0 44.3
UGCB19-147 Bluestone 62.80 76.50 13.70 11.2 78.0
UGCB19-152 Bluestone 39.60 41.90 2.30 49.2 42.0
UGCB19-155 Bluestone 75.30 82.30 7.00 11.9 18.0
UGCB19-157 Bluestone 132.30 139.30 7.00 10.7 131.5
CB19-410 Bluestone 222.40 233.90 11.50 8.5 7.1
CB19-411 Bluestone 215.90 225.40 9.50 7.2 16.0
UGCB20-174 Bluestone 120.83 128.20 7.40 14.9 54.9
UGCB20-176 Bluestone 128.30 142.40 14.10 24.9 38.6
UGCB20-179 Bluestone 61.30 73.10 11.90 86.3 364.9
UGCB20-179 Bluestone 68.60 73.10 4.20 194.0 810.4
CB20-180 Bluestone 170.60 175.93 5.40 334.7 538.8
CB20-181 Bluestone 210.60 215.70 5.10 75.7 32.8
CB20-188 Bluestone 177.70 186.74 9.00 26.0 26.8
CB20-191 Bluestone 24.80 126.20 101.40 2.4 9.6
CB20-420 Bluestone 179.50 195.00 15.50 21.6 51.7
CB20-427 Bluestone 215.80 218.90 3.00 19.1 15.0
CB20-429 Bluestone 22.90 212.14 189.30 0.8 2.5
CB20-430 Bluestone 227.30 236.47 9.30 34.6 66.9
CB20-433 Bluestone 75.60 293.20 217.60 1.4 5.6
CB20-433 Bluestone 293.10 314.30 21.20 11.2 11.7
CB20-442 Bluestone 263.50 292.10 28.60 11.6 12.3
CB20-442 Bluestone 282.60 28.88 6.30 29.0 30.1
CB20-444 Bluestone 54.60 166.30 111.80 2.1 12.5
CB20-444 Bluestone 136.50 143.56 9.50 7.6 55.6
CB20-449 Bluestone 43.30 158.20 114.90 2.5 13.4
CB21-460 Bluestone 114.60 172.21 57.60 3.1 9.9

 

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Hole Company From To Length (m) Au (g/t) Ag (g/t)
CB21-469 Bluestone 1.52 141.73 140.20 1.1 8.2
CB21-487 Bluestone 85.30 92.90 7.60 30.2 85.5

 

Source: Goldcorp, 2014; Bluestone, 2021.

 

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8SAMPLE PREPARATION, ANALYSES AND SECURITY

 

8.1Sampling Method & Approach

 

8.1.1Sampling Preparation, Analyses & Security (prior to November 2006)

 

Prior to Goldcorp taking ownership of the Project in November 2006, all previous drilling, sampling, and assaying were under the control of Glamis.

 

All sample data used in the Era Dorada mineral resource calculations was produced by either diamond drilling (DD) or reverse-circulation (RC) drilling. Drilling contractors were hired to supply the drilling equipment and perform the work under the direct supervision of owner-field personnel.

 

The Glamis drill hole program used a variable combination of sample collection, as follows:

 

·Double-tube HQ core in the upper reaches of the hole switching to double-tube NQ core deeper in the hole.

 

·RC drilling in the upper reaches of the hole above the water table and/or the anticipated mineralization zone, switching to a double-tube NQ core deeper in the hole.

 

·RC drilling for the entire hole.

 

Rotary samples collected from the 4¾ inch, face-sampling, hammer-drilled RC holes were initially collected in a five-gallon bucket. The weight was then recorded, and the sample was placed into the hopper of a Gilson splitter. The process was repeated until the entire 1.5 m sample was collected. The total weight was recorded on the sample sheet along with the sample identification and the time of day collected. Weights were only recorded for the dry portion of the drill hole. The Gilson splitter was set to split the sample into two halves, with one half retained and the other wasted. The remaining 50% was placed into the hopper again, and another 50% split was made. The two samples were placed into pre-labeled plastic sample bags, one for assay and the other for storage. An air hose and nozzle were provided for cleaning the Gilson splitter, pan, and buckets. A geologist was assigned to the rotary rig to supervise sample collection and log geology. A chip tray was created as a permanent record of each hole.

 

The core was collected and placed in wooden core boxes. The core was washed to obtain a clean surface for geological and geotechnical logging and placed in a covered logging facility. All core was photographed on print film. The core was sawn longitudinally with a diamond saw and half the core, on a nominal 1.5 m interval broken at lithologic boundaries, and was placed in pre-labeled plastic bags.

 

The other half was retained for inspection or additional tests as warranted. Splits from the core holes were shipped to a facility operated by CAS Laboratories (CAS Honduras) in Tegucigalpa, Honduras. The unused core was retained for inspection on-site.

 

Samples were transported from Era Dorada to the laboratory in Tegucigalpa, Honduras, by CAS personnel, and all sample preparation and analyses were conducted at CAS Honduras.

 

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Reject samples and pulps were stored at the CAS Honduras facility. Samples were analyzed for gold using a 30 g pulp with a fire assay atomic absorption (AA) finish. Samples that ran over 1.0 g/t Au from this method were re-analyzed for both gold and silver using a 30 g pulp fire assay with gravimetric finish.

 

Glamis had established a limited QA/QC program focused on coarse reject and pulp reject checks. A frequency of 1 in 20 pulps was systematically submitted to the Chemex Laboratories in Nevada for gold and silver analysis in addition to coarse rejects.

 

The drill samples were initially quick-logged to locate and mark significant changes in volcanic stratigraphy. Each volcanic unit was then described, and the location of the structure and their orientations, the percentage of quartz veining, and the type of alteration were recorded.

 

Standard logging conventions were used to capture information from the drill sample. Detailed, daily logging was transcribed onto log sheets and independently entered into Excel spreadsheets. The geologist checked data entry before the data was merged with the main database.

 

Detailed core logging was done by capturing data in four tables: lithology, alteration, Sulfide type, and geotechnical information. Lithology was captured using standardized abbreviations. The alteration was captured as a numeric value corresponding to the alteration type. The visible Sulfide types were captured as a total modal percentage and as relative ratios. Structural data was captured in the “comments/structures” table in the database, as the type and angles taken related to the core axis are displayed in an area as a graphical representation. The geotechnical data recorded rock quality designation (RQD) data for the core portion of the hole.

 

All independent laboratories used in the Project employed quality control procedures and protocols that included duplicates, standard reference materials, and blanks. These were available to Glamis but were not included in assay reports.

 

8.1.2Sample Preparation, Analyses & Security (Goldcorp 2010 through 2012)

 

Drilling completed by Goldcorp (2010 to 2012) was a combination of surface and underground diamond core drilling. Drills were operated by both contract and Goldcorp personnel. The Goldcorp underground drill rig (Boart Longyear LM-75) was used on the surface and converted for underground drilling. Drill holes were developed by drilling a larger diameter (HQ) core at the early stage of the hole, decreasing to NQ and/or BQ if the drilling conditions became difficult.

 

Drill recovery was high, and by utilizing several drill core sizes, Goldcorp was able to ensure drill hole target completion. Drill hole collar surveys were completed using a GPS Trimble system (UTM NAD 27 Zone 16N). In-hole drill surveying for azimuth and dip was completed using the Reflex EZ-Shot system approximately every 50 m along the drill hole.

 

Drill cores (surface and underground) were stored in wooden labeled boxes from the drill and transported to the surface core logging facility at the Era Dorada surface core facility.

 

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Technicians first prepared the core boxes by reviewing drill hole depth tags and reassembling broken sections (from zones of poor recovery).

 

Core logging to identify lithology, alteration, RQD, and sampling selection for core sawing was completed by geologists or technicians under the direction of the geologist. Sampling was also completed by Goldcorp personnel, which included technicians and geologists. The typical sample lengths were 1.0 to 1.5 m with maximum lengths of 2.0 and 3.0 m; sample lengths were based on the lithology and alteration. Logs and the sample database indicated that low-grade and high-grade gold and silver samples were of the same lengths and were not broken out separately or collected in a way that caused sample bias. Samples were collected along the footwall, mineralized zones, and hanging walls without breaks in sampling. Blanks were inserted by Goldcorp personnel when a core sample was submitted. All data was initially collected on paper logs and later transferred to Excel files. This data was then entered in MapInfo™ and MineSight™ software for geological modeling.

 

The core selected for analysis was transported to Inspectorate Laboratories in Guatemala City for sample preparation. Samples were prepared at the Inspectorate (Guatemala) by crushing and pulverizing the drill core to 100 g pulp samples.

 

One pulp sample was sent to Goldcorp’s Marlin Mine for gold assaying (fire assay with AA or gravimetric finish) and silver assaying (AA or AA with gravimetric finish). The second pulp sample was sent to the Inspectorate Laboratory in Reno, Nevada, for gold assaying (fire assay with AA or gravimetric finish) and silver assaying (AA or AA with gravimetric finish). The Marlin Mine assays were completed quickly, which assisted the geologists in developing the drilling program. The Inspectorate assays were used for the purposes of mineral resource modeling and estimation.

 

The QA/QC program employed at the Project was under the direction of Goldcorp. Blank samples were inserted by Goldcorp geologists prior to shipping to the Inspectorate at a frequency of 1 in 25 sample submissions. No duplicates of coarse rejects or standards were included in the QA/QC program at Era Dorada; however, it was recommended that duplicates of the coarse rejects be analyzed and compared and that standards be inserted into the QA/QC sample stream for future drilling campaigns. All analytical results were provided to Goldcorp staff and stored first in Excel and later in MapInfo™ and MineSight™ software. All half-core samples collected by both Goldcorp and Glamis are stored adjacent to the core logging facility on the Project site. The Era Dorada site is fully controlled by perimeter fencing and security. All samples removed from the site were under the control of Inspectorate Laboratories.

 

8.1.3Sampling Preparation, Analyses & Security (Bluestone 2017 to 2021)

 

The drill core from the surface and underground was stored in labeled wooden boxes (Figure 8-1) at the drill site and transported to the surface core logging facility. Before core splitting and logging commence, the drill core was systematically photographed in high resolution using a tripod-mounted camera and digitally archived for reference as part of the drill database.

 

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Figure 8-1: Example of core box photography

 

Source: Bluestone, 2019.

 

Logging and sampling were undertaken on-site at Era Dorada by company personnel under a QA/QC protocol developed by Bluestone. Technicians first prepared the core boxes by reviewing drill hole depth tags, reassembling broken sections, and photographing the core. Core logging to identify lithology, alteration, RQD, and sampling selection for core sawing was completed by technicians under the direction of the geologist. Sampling was also completed by Bluestone technicians. The typical sample lengths are 1.0 to 1.5 m with a minimum sample width of 1 m and maximum lengths of 2.0 m; sample lengths were based on the lithology and alteration. Samples are collected along the footwall, mineralized zones, and hanging walls without breaks in sampling. All data was initially captured on paper logs and later transferred to Microsoft Excel. The data was then entered into MapInfo™ and MineSight™ software for geological modeling.

 

Specific gravity readings of all representative lithologies and vein material were taken during the various drill campaigns using the displaced water method. Samples were sealed with paraffin wax to account for natural voids/vugs.

 

A total of 591 channel samples were taken along representative veins exposed in the side walls of the Era Dorada underground tunnels using a portable rock saw. The sampling was undertaken across and perpendicular to the mineralized structures wherever possible and carefully surveyed with XYZ coordinates for use in 3D modeling. The samples were subject to the same QA/QC protocols as the drill core and were deemed suitable for use in calculating resources. Figure 8-2 shows a saw-cut channel sample across a mineralized vein in the South Ramp of the Era Dorada underground workings.

 

 

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Figure 8-2: Example of underground channel sample

 

Source: Bluestone, 2019.

 

Samples were transported in security-sealed bags to Inspectorate Laboratories in Guatemala City for sample preparation until March 2020 and thereafter to Inspectorate Laboratories in Managua due to the closure of the Guatemalan facility. Samples were prepared at the Inspectorate by crushing and pulverizing the drill core down to 85%, passing -75 µm. Pulps were weighed and individually packaged into 100 g envelopes and shipped for analysis. Both coarse rejects and pulp were stored for future use and utilized in Bluestone’s QA/QC program. All half-core and coarse rejects are stored adjacent to the core logging facility on the Project site. The Era Dorada site is fully controlled by perimeter fencing and security.

 

Pulps are shipped for regular and QA/QC analysis to Inspectorate Laboratories (a division of Bureau Veritas) in Reno, Nevada, USA, and ALS Chemex in Vancouver, BC, Canada, respectively. Both are ISO 17025-accredited laboratories. Gold and silver were analyzed by a 30 g charge with atomic absorption with gravimetric finish for values exceeding 5 g Au/t and 100 g Ag/t.

 

All analytical results were provided to Bluestone by respective laboratory secure servers in Excel, .csv, and .pdf formats (certificates). Bluestone database files are stored and managed in Access and Excel formats before being transferred to MapInfoTM and MineSightTM software.

 

During Q3 and Q4 2020, the Cerro Blanco database was transitioned to the AcQuire/GMSuite platform, providing an enhanced, secure, and high standard of data management.

 

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8.2Quality Assurance & Quality Control

 

8.2.1QA/QC Performance & Discussion for Samples prior to 2017

 

Field blanks of non-mineralized material were inserted into the sample series every 25 samples (4%) to test for any potential carry-over contamination that might occur in the crushing phase of sample preparation due to poor cleaning practices. A total of 1,390 blanks were analyzed, with 558 performed at Inspectorate Laboratories, 302 at CAS Honduras, and 530 at the Marlin Mine laboratory. An analysis of the Inspectorate blanks resulted in five fails or 0.01%, with one re-failing on resample. This appears to be the result of sample misclassification as both the original and resample are relatively high grade. The CAS Honduras results showed eight fails or 0.03%, with four of those failing on resample. There may have been some cleaning issues at CAS Honduras, although it was not widespread or significant. The blanks from the Marlin Mine laboratory resulted in 14 fails or 0.03%, which is not significant. Considering that the Marlin Mine assaying was utilized for fast turnaround to guide the program and not for resource estimation purposes, this fail rate does not pose an issue.

 

Core duplicate samples were used to evaluate analytical precision and to determine if any biases exist between laboratories that may affect the overall assay database. The core duplicate samples were quarter-spilt cores sampled on-site and sent to Inspectorate Laboratories and CAS Honduras. A total of 1,060 samples with gold values >2 g/t were selected in the drill hole database through hole CB-222. Of those, a total of 797 samples were submitted for check analyses, with 618 samples being submitted to the Inspectorate for checks of original CAS Honduras analyses, while 179 samples were submitted to CAS Honduras for checks of original Inspectorate analyses. The 618 Inspectorate duplicate check samples show the CAS Honduras original samples to be 3% higher in gold and 16% higher in silver on an individual basis and 3% and 2.8% higher in gold and silver, respectively, on an overall basis.

 

The 179 CAS Honduras duplicate check samples show the Inspectorate original samples to be 1.5% lower in gold and 27% lower in silver on an individual basis and 6.8% and 11.4% lower in gold and silver, respectively, on an overall basis.

 

Duplicate analyses from both labs show high variation in individual gold values, potentially attributable to the nugget effect, particularly for higher-grade samples. However, on average, the samples show a better correlation, which has greater implications on a global or resource scale. The CAS Honduras check samples appeared to show a relatively small grade bias.

 

Standards are used to test the accuracy of the assays and to monitor the consistency of the laboratory over time. Neither Glamis nor Goldcorp employed the use of standards. It was recommended that a QA/QC program be implemented during all future drill programs that include the insertion and analysis of standards, blanks, and duplicates, as well as umpire assays.

 

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8.2.2QA/QC Performance & Discussion of Results (Bluestone 2017 to 2021)

 

Since 2017, Bluestone has implemented a comprehensive QA/QC program employing industry standards and best practices for all its drill core and channel sampling. This includes the insertion of blind-certified reference materials (blanks and standards) into the sample stream, in addition to field blanks. Furthermore, duplicate analysis of pulps and coarse rejects was performed at a second laboratory to independently assess the analytical precision and accuracy of each sample batch as they were received from the laboratory. Additionally, pulp and coarse rejects were systematically submitted to ALS Chemex Laboratories in Vancouver for check analysis and additional quality control.

 

A total of 7,652 control samples (Table 8-1) were assigned for QA/QC purposes, accounting for approximately 20% of the total samples taken during the program.

 

Table 8-1: Quantity of control samples by type (Bluestone 2017 to 2021)

 

Control Type Number
Standards 1,602
Field Blanks 685
Pulp Blanks 859
Pulp and Coarse Reject Duplicates 4,506
Total 7,652

 

Source: Bluestone, 2021.

 

Standards are used to test the accuracy of the assays and to monitor the consistency of the laboratory over time. A variety of certified standards of various gold grades were purchased from CDN Laboratories (Table 8-2) and inserted by the logging geologists.

 

Table 8-2: Summary of standards (Bluestone 2017 to 2021)

 

Control Sample Au PPM Standard Deviation Analysis
CDN-GS-16 16.48 0.315 Fire Assay Gravimetric
CDN-GS-11B 11.04 0.44 Fire Assay Gravimetric
CDN-GS-6F 6.79 0.15 Fire Assay Gravimetric
CDN-GS-6E 6.06 0.16 Fire Assay Gravimetric
CDN-GS-5T 4.76 0.105 Fire Assay AA Finish
CDN-GS-1W 1.063 0.038 Fire Assay AA Finish
CDN-GS-1T 1.08 0.05 Fire Assay AA Finish
CDN-GS-1X 1.299 0.06 Fire Assay AA Finish
CDN-BL-10 <0.01 - Fire Assay AA Finish
FIELD BLANKS <0.01 - Fire Assay AA Finish

 

Source: Bluestone, 2021.

 

Field blanks are non-mineralized materials sourced locally that are inserted into the sample series every 20 samples (5%). Field blanks are inserted to test for any potential carry-over contamination that might occur in the crushing phase of sample preparation due to poor laboratory cleaning practices.

 

Duplicate analysis of pulps and quarter-core are used to evaluate the analytical precision and to determine if any biases exist between laboratories. Duplicate analysis of coarse rejects is used to analyze preparation errors. Table 8-3 shows the QA/QC sample insertion rate.

 

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QA/QC assay results were checked by a Bluestone database QA/QC manager on a batch-by-batch basis for analytical or batch errors. No evidence of obvious analytical bias was noted. Figure 8-3 shows a control plot for standard CDN-GS-6E.

 

Table 8-3: Bluestone QA/QC sample insertion rates

 

Batch Size – 45
Samples
Minimum
Insertion Rates
Notes
Standards 1 every 20 Inserted according to the estimated grade of mineralization before, within, or immediately after a mineralized interval. Insertion at regular intervals avoided.
Field Blanks 1 every 20 Usually inserted at the end of mineralized runs to measure carry-over
Pulp Blanks 1 every 20 Usually inserted at the end of mineralized runs to measure carry-over
Pulp Duplicates 1 every 20 Undertaken at the second laboratory with the same analytical technique. High- and low-grade mineralized samples are usually chosen
Coarse Duplicates 1 every 20 Normally choose mineralized samples, used to measure laboratory sample preparation

 

Source: Bluestone, 2020.

 

 

 

Figure 8-3: Batch plot of standard CDN-GS-6E

 

Source: Bluestone, 2020.

 

Except for one standard, the performance of the control samples was very good, reflecting the overall high quality of the analysis. Standard CDN-GS5T (4.76 g Au/t) utilized early in the Bluestone drill program plotted consistently along the highest acceptable threshold for fire assay with instrumental finish. Check analysis at both the Inspectorate and ALS Chemex laboratories gave similar results. As lower-grade CRM / blanks and the laboratories’ internal QA/QC procedures ruled out any calibration issues, the use of this particular standard was discontinued.

 

Duplicates of pulp and coarse rejects were sent to ALS Chemex in Vancouver for check gold analysis with the analysis at the principal laboratory, Inspectorate Laboratories in Reno. As shown in Figure 8-4, the results indicate a very good correlation at both low and high gold levels and excellent reproducibility between the two laboratories, with a correlation coefficient of 0.993.

 

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The results can be interpreted as a reflection of the micron-sized nature of the gold and the lack of coarse, nuggety gold in the Era Dorada deposit. Analyses of both pulp and field blanks (Figure 8-5) consistently yielded gold values near or below the detection limit of the primary laboratory. No sample contamination was detected.

 

 

 

Figure 8-4: Plot of pulp & coarse reject duplicates (Bluestone 2017-2021)

 

Source: Bluestone, 2021.

 

 

 

Figure 8-5: Pulp & field blanks (Bluestone 2017 to 2021)

 

Source: Bluestone, 2021.

 

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It is the opinion of the QP, Garth Kirkham, P. Geo., that the sampling preparation, security, analytical procedures, and quality control protocols used by Bluestone are consistent with generally accepted industry best practices and are, therefore, reliable for the purpose of resource estimation.

 

The Qualified Person is of the opinion that the sample preparation, security, and analytical procedures are adequate for the purpose of mineral resource estimation as presented within this Technical Report Summary.

 

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9DATA VERIFICATION

 

A geological site visit is a critical part of the due diligence process that ensures mineral disclosures are accurate, independently verified, and based on sound technical observations. Multiple site visits were conducted by several of the QP, as detailed in Section 2.2. The purpose of a site visit in the context of Securities and Exchange Commission Regulation S-K Subpart 1300 is to provide qualified third-party review and verification of the geological, technical, and operational aspects of a mineral property. These site visits consisted of underground investigations of mineralized and non-mineralized headings, as well as an inspection of the surface core logging, sampling, storage areas, and existing infrastructure.

 

The QP performed an independent verification of the data, observations, and interpretations for Era Dorada. This included confirmation sampling procedures, drilling methods, core logging, and QA/QC practices. Inspected drill cores, outcrops, underground workings, and surface trenches to corroborate reported geological models, historical data, and reporting. Additionally, this involved a thorough examination of mining infrastructure access, along with an extensive review of environmental and social conditions. Identification and evaluation of risk in support of the mineral resource/ estimates.

 

9.1Geology, Drilling & Assaying

 

Garth Kirkham, P. Geo., has been involved with the property since its acquisition in early 2017, when he performed the initial due diligence and authored the updated resource estimate for Bluestone. Mr. Kirkham first visited the property on May 8, 2017, to validate all aspects. The site visit included an inspection of the property, offices, underground vein exposures, core storage facilities, water treatment plant, and stockpiles, and a tour of major centers and the surrounding villages most likely to be affected by any potential mining operation.

 

Since 2017, Mr. Kirkham has visited the property numerous times for extended periods to develop and implement data gathering and sampling methods and procedures. He also worked with Bluestone geologists to develop drill programs and to supervise interpretation and modeling efforts in addition to creating and implementing QA/QC procedures.

 

From September 21 to 22, 2017, Mr. Kirkham inspected the progress of the recommended historic drill core rehabilitation program and initiated structural studies.

 

From April 24 to 28, 2018, Mr. Kirkham’s site visit focused on advancing the planning of sampling and drilling along with supporting lithological and structural modeling.

 

From February 16 to 22, 2020, Mr. Kirkham provided guidance on the planning and development of advanced drilling and sampling, as well as grade vein modeling.

 

From January 10 to 15, 2021, Mr. Kirkham assisted with validating drill and sample data, refining high-grade models, reviewing low-grade models, and providing guidance for the finalization of the open pit bulk tonnage resource scenario.

 

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Continued data validation and verification processes have not identified any material issues with the Era Dorada sample and assay data. Mr. Kirkham is satisfied that the assay data is of suitable quality to be used as the basis for this resource estimate.

 

During Q3 and Q4 2020, the Era Dorada drill and assay database was switched over to the AcQuire - GMSuite platform hosted by CSA Global, providing an enhanced and more secure standard of data management.

 

Mr. Kirkham is confident that the data and results are valid and can be relied upon. Mr. Kirkham is also confident that the methods and procedures used are reliable. It is the opinion of Mr. Kirkham that all work, procedures, and results have adhered to best practices and industry standards.

 

The Qualified Person is of the opinion that the data is adequate for the purposes used within this Technical Report Summary.

 

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June, 2025
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10MINERAL PROCESSING AND METALLURGICAL TESTING

 

10.1Introduction

 

Various metallurgical testing campaigns were conducted on Era Dorada (formerly named “Cerro Blanco”) samples by Kappes, Cassiday & Associates (KCA) between 1999 and 2012, together with additional testing carried out by SGS Lakefield Research Ltd., Carson GeoMIn Inc., Pocock Industrial Inc., Phillips Enterprises Inc., and CyPlus GmbH. The most recent test program was completed in 2018 and was carried out at Base Metallurgical Laboratories Ltd. (BaseMet) in Kamloops, BC. The following reports include all metallurgical testing programs carried out so far on Era Dorada samples.

 

·Kappes, Cassiday & Associates (1999): “Cerro Blanco Project, Results of Cyanide Leach Tests” (Issued: April 8, 1999).

 

·Kappes, Cassiday & Associates (2000a): “Cerro Blanco Project, Results of Cyanide Bottle Roll Tests” (Issued: January 12, 2000).

 

·Kappes, Cassiday & Associates (2000b): “Cerro Blanco Project, Bottle Roll Tests” (Issued: August 24, 2000).

 

·Kappes, Cassiday & Associates (2002): “Cerro Blanco Project, Results of Leaching Tests and Gravity Concentration Tests” (Issued: 8 October 2002).

 

·SGS Lakefield Research Ltd. (2005): “Cerro Blanco North Zone Samples for Met Testing at SGS Lakefield” (Issued: August 2005).

 

·Kappes, Cassiday & Associates (2005): “Cerro Blanco Project” (Issued: December 15, 2005).

 

·Carson GeoMIn Inc. (2005): “Mineralogy of Ore Composites and Related Cyanide Tailings from the Cerro Blanco Gold Project” (Issued: December 29, 2005).

 

·Kappes, Cassiday & Associates (2006a): “Cerro Blanco Project” (Issued: January 18, 2006).

 

·Kappes, Cassiday & Associates (2006b): “Cerro Blanco Project” (Issued: April 21, 2006).

 

·Phillips Enterprises LLC (2011): “Comminution Tests, Cerro Blanco” (Issued: June 11, 2011).

 

·Pocock Industrial Inc. (2011): “Sample Characterization and PSA, Flocculant Screening, Gravity Sedimentation, Pulp Rheology, Vacuum Filtration and Pressure Filtration Studies Conducted for Kappes, Cassiday & Associates Cerro Blanco Project” (Issued: October 2011).

 

·Kappes, Cassiday & Associates (2012): “Cerro Blanco Project, Report of Metallurgical Test Work, January 2012” (Issued: January 25, 2012).

 

·Base Metallurgical Laboratories Ltd. (2018): “BL0246: Generation of Cyanide Detox Tailings – Cerro Blanco Project” (Issued: August 3, 2018).

 

The following sections show the selected reports and respective testing results for designing the recovery method and equipment for the Project.

 

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10.2Selected Testworks

 

10.2.1KCA (2012) – Leach Tests

 

The six pallets received by KCA in April 2011 included a total of 55 cloth bags containing half-split HQ and PQ drilling core material from five samples. These five samples were assayed, together with a master composite sample. The latter was assembled on the basis of the same five samples. The head assay results of both Au and Ag are summarized in Table 10-1.

 

Table 10-1: Head assays

 

KCA Sample
No.
Description Average Au Assay
(g/t)
Average Ag Assay
(g/t)
48901 MbT 9.4 46.25
48902 Mcv 4.47 5.79
48903 Svc 6.52 44.71
48904 Msc 5.07 38.79
48905 Cbx 4.59 18.06
48907 Master Composite 7.7 37.86

 

Source: KCA, 2012.

 

The experimental program included bottle roll leach testing on the Master Composite sample according to different grinding sizes. The summarized testing conditions and gold extraction results are listed in Table 10-2, which indicates that for P80 (80% passing) equal to or finer than 0.085 mm, the overall gold extractions ranged from 92% to 94%.

 

Table 10-2: Gold extraction summary

 

KCA Sample
No.
KCA Test
No.
P80 – Milled Size
(µm)
Target NaCN
(g/l)
Calculated Head –
Au (g/t)
Gold Extracted –
Au (%)
Leach Time
(h)
48907 48914 A 160 1.0 6,537 86 96
48907 48913 A 138 1.0 6,328 91 96
48907 48913 B 85 1.0 5,822 93 96
48907 48913 C 77 1.0 6,497 94 96
48907 48914 B 76 1.0 5,887 92 96
48907 48913 D 72 1.0 5,628 94 96
48907 48916 C 71 1.0 5,788 92 96
48907 48915 B 57 1.0 6,124 93 96
48907 48916 B 55 1.0 8,663 94 96
48907 48917 B 44 1.0 5,169 92 96
Average 1.0 6,244 92.1 96

 

Source: KCA, 2012.

 

10.2.2Phillips Enterprises (2011) – Comminution Tests

 

The same five samples used in the KCA testing campaign listed in Section 10.2.1 were split and used for comminution testing at Phillips Enterprises LLC in Golden, Colorado. Comminution testing included the Bond Work index for the ball mill (BWi), Bond Work index for the rod mill (RWi), as well as Bond Abrasion index (Ai). The results obtained are summarized in Table 10-3.

 

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Table 10-3: Comminution test results

 

KCA
Sample No.
Sample
Description
Bond Rod Mill Work
Index - BWi (kWh/t)
Bond Ball Mill Work
Index - RWi (kWh/t)
Bond Abrasion
Index - BAi
48901 MbT 17.08 20.27 0.193
48902 Mcv 13.91 16.37 0.104
48903 Svc 18.26 22.24 0.328
48904 Msc 16.90 21.45 0.329
48905 Cbx 15.52 18.95 0.246
Average 16.33 19.86 0.240

 

Source: Phillips Enterprises, 2011.

 

10.2.3Pocock Industrial (2011) – Solid / Liquid Separation Tests

 

Solid/liquid separation tests were conducted on “Fresh Milled” and “Leached and Detoxed” samples as resulting from the KCA test work previously described in Section 10.2.1. The program included testing for supporting solid/liquid separation equipment design and sizing. All testing was conducted by Pocock Industrial at their laboratory facilities located in Salt Lake City, Utah during October 2011.

 

The summary of solid/liquid separation testing was as follows.

 

·The flocculant concentration varied by individual sample and thickener type or application but were in the overall range of 35 to 55 g/t for Fresh Milled sample, as well as 30 to 55 g/t for Leached and Detoxed sample within the tested pH range.

 

·For conventional thickener sizing, Pocock recommended a minimum unit area design basis of 0.30 to 0.40 m²/tpd for the Fresh Milled, and 0.25 to 0.35 m²/tpd for the Leached and Detoxed material.

 

·Dynamic thickening tests conducted on the samples indicated a hydraulic net feed loading rate design basis in the maximum range of 3.1–4.3 m³/m²·hr for both the Fresh Milled and Leached and Detoxed samples for high performance.

 

·The overall maximum underflow density range for the Leached and Detoxed material was 53 to 57%. However, the range was narrowed to 53 to 55% with rake torque considerations based on un-sheared data.

 

·The designed pressure filter for a stipulated 1,250 tpd plant throughput resulted in for both Leached and Detoxed samples a minimum requirement of 190 chambers for a horizontal recess plate type press, equipped with 1,500 mm plates and 15 mm recess (30 mm full chamber) with no cake wash to achieve 18.3% moisture. Alternative design indicated equipment with 181 chambers to achieve 18.9% moisture with pH adjusted to 10.5.

 

10.2.4BaseMet (2018) – Chemical Assays

 

The 2018 testing campaign carried out at BaseMet was based on the two following samples.

 

·Approximately 90 kg of half and quarter cut drill core.

 

·About 590 kg of bulk rock, consisting of 180 individual intervals.

 

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Each drill core interval was stage crushed to 3.36 mm. The crushed material was blended and split to obtain a representative sub-sample of the Global Composite. The head assay results are shown in Table 10-4.

 

Table 10-4: Head assays

 

Composite Au (g/t) Ag (g/t) Cu (%)
Global Composite Head 1 4.21 23 0.007
Global Composite Head 2 5.65 21 0.007
Average 4.93 22 0.007

Source: BaseMet, 2019.

 

10.2.5BaseMet (2018) – Gravity Concentration

 

The same testing campaign described in Section 10.2.4 comprised of gravity concentration testing. Accordingly, sub-samples of the Global Composite sample were ground in a laboratory rod mill to three targeted P80 grind sizes of 0.050 mm, 0.075 mm and 0.100 mm. Each one of these three samples were tested at a Knelson MD-3 centrifugal gravity concentrator. Knelson concentrates were panned targeting a 0.1% to 0.5% mass recovery. Gravity concentration results shown in Table 10-5 indicate average recoveries of 19.1% Au and 7.8% Ag.

 

Table 10-5: Gravity concentration results

 

Test No. Grind Size (µm) Mass Recovery (%) Au Recovery (%) Ag Recovery (%)
4 50 0.317 22.5 6.3
10 50 0.186 21.1 6.9
11 50 0.23 15.1 6.5
17 53 0.319 20.8 9.4
18 53 0.301 17.9 8.5
19 53 0.274 16.7 6
20 53 0.326 21.7 9.1
21 75 0.185 16.3 5.4
2 75 0.239 29.8 16.5
6 75 0.27 14.7 5.4
7 75 0.314 20.7 6.4
8 75 0.398 20 6.6
3 75 0.48 17.7 10.2
12 75 0.291 15.9 4.1
5 100 0.534 15.6 10.2
Average 0.311 19.1 7.8

Source: BaseMet, 2019.

 

10.2.6BaseMet (2018) – Leach Tests

 

Leaching test work was also carried out in the 2018 BaseMet testing campaign. In this case, the tests consisted of direct cyanide leaching on two samples, i.e., fresh milled product and gravity tailings. All tests were conducted in closed rolling bottles with monitoring and controlling of cyanide level, dissolved oxygen (DO), and pH. Sampling for kinetic assessments was conducted during test periods of 2, 6, 24, 48, and 72 hours. The leach test results are summarized in Table 10-6. The average recoveries were 93.24% Au for 24 hours of residence time and 94.96% Au for 72 hours of residence time.

 

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Table 10-6: Bottle roll leach results

 

Test No. Grind
Size (µm)
Reagent Consumption Gravity Au
Recovery (%)

Cumulative Gold

Extraction (%)

Final Gold
Recovery (%)
Final Ag
Recovery (%)
NaCN (kg/t) Lime (kg/t) 2 h 6 h 24 h 48 h 72 h 72 h
4 50 0.84 0.87 22.5 75.3 92.4 95.9 95.7 96.1 92.4
10 50 0.36 1.17 21.1 78.9 92.4 94.4 95.7 97.5 69.6
11 50 0.52 1.02 15.1 81.5 92.7 94.0 96.5 97.3 78.6
17 53 0.52 1.22 20.8 91.9 93.2 94.2 95.2 95.9 88.4
18 53 0.50 1.12 17.9 82.2 91.6 96.6 95.1 96.1 92.3
19 53 0.86 1.33 16.7 87.7 94.9 98.1 97.4 94.5 70.8
20 53 0.60 1.50 21.7 80.7 90.9 94.1 92.8 96.3 69.7
21 53 0.28 0.96 16.3 80.9 90.1 91.5 91.6 94.7 67.2
2 75 0.82 0.86 29.8 61.2 78.5 91.9 94.1 94.7 86.9
6 75 0.82 0.84 14.7 82.9 90.4 92.4 93.1 94.2 82.9
7 75 1.00 0.71 20.7 77.8 90.8 92.9 93.7 94.4 84.2
8 75 0.46 0.82 20.0 75.0 88.3 92.7 93.2 93.6 83.1
3 75 0.76 0.89 17.7 68.6 87.2 92.5 93.0 94.0 93.2
12 75 0.20 1.00 15.9 80.6 89.3 91.7 93.8 95.6 65.2
5 100 0.58 0.71 15.6 66.3 82.4 91.2 91.7 91.9 82.7
1 75 2.98 0.50 No Gravity 3.2 10.1 88.4 92.2 93.1 84.7
9 75 0.90 0.77 No Gravity 67.4 86.5 92.6 92.1 94.4 86.3
Average 0.76 0.96 19.10 73.06 84.81 93.24 93.94 94.96 81.07

Source: BaseMet, 2019.

 

According to the results listed in Table 10-6, finer grinding sizes resulted in higher Au and Ag recoveries.

 

Figure 10-1 shows a graph of gold extraction as a function of the leaching period (residence time) for different grind sizes. The plotted curves include gravity recovery. The average gold extraction for a P80 grind size of 0.050-0.053 mm was 95% for 24 hours, as well as 97% for 72 hours.

 

 

Figure 10-1: Effect of grind size on average gold extraction

 

Source: Author; BaseMet, 2019.

 

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One of the Composite samples previously listed in Section 10.2.4 was further tested to determine the adsorption of Au and Ag on carbon and, therefore, the basis for a future Carbon in Pulp (CIP) circuit. The conditions for CIP testing were as follows:

 

·Sample grinding: P80 of 0.053 mm.

 

·Pulp Density: 33% solids (w/w).

 

·Pulp pH: 10.5 to 11 maintained with lime.

 

·Cyanide Concentration: 0.5 g/l NaCN maintained.

 

·Retention Time: Total 54 h (48 h leach, 6 h carbon absorption).

 

·Carbon Concentration: 25–50 g/l.

 

·Lead Nitrate Concentration: 0–250 g/t.

 

The recoveries obtained for Au and Ag are indicated in Table 10-7. Tests carried out at 50 g/l carbon concentration (T-25, T-26, and T-27) resulted in higher Au and Ag recoveries. It was also observed that the addition of lead nitrate improved Ag recovery (T-26 and T-27).

 

Table 10-7: Bottle roll leach results (CIP)

 

Test
ID
Grinding
P80
Total Leaching
Time
NaCN Consumption
PbNO3
Carbon
Concentration
Au
Calculated
Gravity
Au
Extraction
(%)
(µm) (h) (g/l) (kg/t) (g/l) (g/t) (%) Au Ag
T-21 53 54 0.5 0 25 6.95 16.3 94.7 67.2
T-25 53 54 0.5 0 50 6.95 33.2 97.1 81.1
T-26 53 54 0.5 250 50 6.64 28.6 97.3 90.0
T-27 53 54 0.5 250 50 6.59 25 96.6 85.4

Source: BaseMet, 2019.

 

The Au recovery figures listed in Table 10-7 were plotted in a graph, as shown in Figure 10-2. Overall, Au recovery for a 36-hour leaching period was estimated as 96% (Tests T-25, T-26, and T-27).

 

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Figure 10-2: Gold recovery as a function of residence time (CIP)

 

Source: BaseMet, 2019.

 

10.2.7BaseMet (2018) – Cyanide Destruction Tests

 

A series of continuous cyanide destruction tests were conducted to assess the efficiency of the selected detox process for the leaching test tailings.

 

Samples from leach tests described in Section 10.2.6 were filtered, and the resulting solution was analyzed. The results indicated 283 mg/l total cyanide (CNT) and 270 mg/l weak acid dissociable cyanide (CNWAD).

 

The SO2/air method was assessed for destructing cyanide contained in leached tailings. Such a method, also known as the Inco or Detox method, uses sulfur dioxide (SO2) to remove weak acidic dissociable cyanide down to concentrations smaller than 5 mg/l.

 

The cyanide pulp resulting from leach tests performed adequately to the SO2/Air cyanide destruction process, resulting in final tailings with less than 1 mg/l CNWAD, as well as less than 4 mg/l total cyanide (CNT). The results shown in Table 10-8 indicate that the Detox tests resulted in CNWAD concentrations smaller than 5 mg/l in the final tailings.

 

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Table 10-8: Cyanide destruction results

 

Test No. Retention
Time (min)
pH Reagent Addition (g / g CNWAD) Cu (mg/l of
solution)
Final Solution Composition
SO2 Equivalent Lime CNT (mg/l) CNWAD (mg/l)
CND-C1 90 8.5 7 4.6 100 4.59 1.8
CND-C2 90 8.5 5.5 2 100 2.96 0.17
CND-C3 90 8.3 4 2.2 100 0.49 0.22
CND-C4 90 8.4 4 1.6 50 2.94 0.14
CND-C5 90 8.5 4 1.4 25 3.02 0.24
CND-C6 90 9 4 - 0 18.3 4.18
CND-C7 60 8.5 4 0.8 25 3.56 0.48

Source: BaseMet, 2019.

 

10.3Summary and Conclusions

 

The summary and main conclusions of the selected test work carried out on Era Dorada samples were as follows.

 

·Comminution: Bond Ball Mill Work index (BWi) of 19.9 kWh/t and a Bond Abrasion index (Bai) of 0.24.

 

·Grinding size: P80 of 0.053 mm.

 

·Gravity Concentration: Based on the gravity test results, a gravity concentration circuit was included in the grinding circuit.

 

·Leach Results: Oxygen pre-oxidation should be incorporated into the process design, as a significant impact was observed during the first 24 hours of leaching. The recommended residence time of the pre-oxidation stage was 2 hours at a targeted cyanide concentration of 500 ppm.

 

·Leach Results (CIP): The results and test work parameters obtained from the four leach tests were used to develop the process design criteria and estimated Au and Ag recoveries for the leach/CIP circuits.

 

·Cyanide Destruction: The conditions used in tests CND-C7 were adopted as process design for the cyanide destruction circuit for reducing the CNWAD concentration to less than 1 mg/l.

 

·Leach Results: The relatively small increase in Au recovery for a leaching period of 72-hour as compared with the 36-hour leaching period resulted in the adoption of the latter for the Era Dorada industrial plant. Detailed figures are listed in Table 10-9. The leached slurry will then flow through a carbon-in-pulp (CIP) circuit for the adsorption of the Au and Ag cyanide complexes onto the pores of activate carbon. The loaded carbon will be processed through desorption and refining circuits. The adopted Au and Ag recoveries should be 96% and 85% respectively.

 

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Table 10-9: Preliminary recovery estimative

 

Test ID Residence Time 54 h Residence Time 36 h  
 
Recovery (%) Estimated Recovery (%)  
 
Au Ag Au  
T-25 97.1 81.1 95.8  
T-26 97.3 90.0 95.8  
T-27 96.6 85.4 95.4  
Average 97 86 96  

Source: Author; BaseMet, 2019.

 

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11MINERAL RESOURCE ESTIMATES

 

11.1Introduction

 

This section describes the work undertaken by Kirkham Geosystems Ltd (KGL), including key assumptions and parameters used to prepare the mineral resource models for Era Dorada, together with appropriate commentary regarding the merits and possible limitations of such assumptions.

 

Era Dorada is a classic hot springs-related, low-sulfidation epithermal gold-silver deposit comprising both high-grade vein and low-grade disseminated mineralization. Most of the high-grade mineralization is hosted in the Mita unit as two upward-flaring vein swarms (north and south zones) that converge downwards and merge into basal feeder veins where drilling has demonstrated widths of high-grade mineralization (e.g., 15.5 m 21.4 g Au/t and 52 g Ag/t). Bonanza gold grades are associated with ginguru banding and carbonate replacement textures. Sulfide contents are low, typically < 3 volume %.

 

The Mita rocks are overlain by the Salinas unit, a sub-horizontal sequence of volcanogenic sediments and sinter horizons approximately 100 m thick that form the low-lying hill at the project. Low-grade disseminated and veinlet mineralization within and as halos around the high-grade vein swarms is well documented in drilling since discovery of the deposit, with grades typically ranging from 0.3 to 1.5 g Au/t. The overlying Salinas cap rocks are also host to low-grade mineralization associated with silicified conglomerates and rhyolite intrusion breccias.

 

The mineral resource has a footprint of 800 x 400 m between elevations of 525 and 200 m above sea level (masl). The mineral resource estimate is the result of 141,969 m of drilling by Bluestone and previous operators (1,256 drill holes and channel samples by Bluestone). The 3.4 km of underground infrastructure allowed for underground mapping, sampling, and over 30,000 m of underground drilling that enhanced the current understanding and validation of the Era Dorada geological model. The mineral resource estimate is based on a scenario that considers open pit mining methods and therefore requires improved and refined geological models of the lithologic units. These broad mineralised lithologies are host to the high-grade veins that have been the focus of the potential underground mining scenario. The resulting domain models and estimation strategy were designed to accurately represent the grade distribution.

 

Several resource estimates have been published on Era Dorada since 2017 in four technical reports, as follows:

 

·Preliminary Economic Assessment (March 20, 2017)

 

·Preliminary Economic Assessment Update (June 2, 2017)

 

·Feasibility Study (January 29, 2019)

 

·Preliminary Economic Assessment Update (February 28, 2021)

 

·Preliminary Economic Assessment Update (June 30, 2021)

 

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The first three reports and resource estimates were for an underground mining scenario. The last two resource estimates were for the open pit scenario. All estimates were authored by Qualified Person, Garth Kirkham, P.Geo.

 

All five technical reports are filed on the System for Electronic Document Analysis and Retrieval (SEDAR+).

 

11.2Data

 

The drill hole database was supplied in electronic format (i.e., Microsoft Excel and Access) by Bluestone. This included collars, down hole surveys, lithology data and assay data (i.e., grams per tonne of gold and silver, and down hole “from” and “to” intervals in metric units). Lithology group and description information was provided, along with abbreviated alpha-numeric and numeric codes (see Table 11-1). Figure 11-1 shows the plan view of drill holes with collars. A total of 130,238 assay values and 55,285 lithology values were supplied for the project. Validation and verification checks were performed during import to confirm there were no overlapping intervals, typographic errors, or anomalous entries.

 

Table 11-1: Lithology units & codes

 

Lithology Code Code B Lithology Group Lithology Description
Qc 10 1 Post-Mineral Cover Rock - Quaternary Colluvium
Qb 11 1.1   Basalt Flows
Bi 20 2 Cross-Cutting Rock Types Basaltic Intrusive Dikes
Cbx 30 3   Collapse Breccia
Dp 180 18   Dacite
Gr 40 4   Granite
Ad 50 5   Andesite Dike
Rp 60 6   Quartz Eye Rhyolite
Vt 70 7   Vein
Stock 71 7.1   Stockwork
Hbx 72 7.2   Hydrothermal Breccia
RF 80 8   Rhyolite Flow
SZ 81 8.1   Shear Zone
Ss 90 9 Salinas Group Sinter
Svc 91 9.1   Volcanic Sediments
Srt 92 9.2   Quartz Eye Rhyolite
Sfx 93 9.3   Phreatic Breccia
Slt 94 9.4   Siltstone
Sct 95 9.5   Ash Tuff
Scgl 96 9.6   Conglomerate
Mss 100 10 Mita Group Sandstone
Mat 101 10.1   Andesite Tuff
Mlt 102 10.2   Crystal Tuff
Mbt 103 10.3   Lapilli Tuff
Msc 104 10.4   Calcareous Limestone
Mls 105 10.5   Limestone
Mcv 106 10.6   Quartz Latite Crystal Lithic Tuff
Mvo 107 10.7   Conglomerate
Mlm 190 19   Upper Limestone
Silt 108 10.8   Siltstone - mudstone
PA 130 13   Porphyritic andesite
Tcb 110 11 Tempisque Volcanic Complex Basalt-dominated

 

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Lithology Code Code B Lithology Group Lithology Description
Tca 111 11.1   Andesite-dominated

Source: Kirkham, 2025.

 

 

Figure 11-1: Plan view of drill holes

 

Source: Kirkham, 2025.

 

11.3Data Analysis

 

Table 11-2 shows statistics of gold and silver assays for each of the lithologic units. It should be noted that the total number of values from section to section vary depending on the parameter being analysed and the value for reporting these varied data sub-sets is to detect and investigate issues or anomalies. Included for the statistical analysis, there are 130,307 gold assays (153,078 m) total, which average 0.68 g/t, and there are 130,238 (153,003 m) silver assays by lithology logged, which average 3.75 g/t. The maximum gold assay is 1,380 g/t, while the maximum silver assay is 8,656.7 g/t. It is important to note that 73 gold assays are greater than 100 g/t and 54 silver assays are greater than 500 g/t which may be a reflection of the non-nuggety nature of the mineralization present at Era Dorada.

 

Table 11-2: Statistics for weighted gold & silver assays

 

Code Metal Valid Length (m) Max (gpt) Mean (gpt) CV
Total AU 130,307 153,077.8 1,380.0 0.68 9.9
AG 130,238 153,003.0 8,656.7 3.75 11.1
All AU 131,215 154,481.6 1,380.0 0.69 9.8
AG 131,146 154,406.9 8,656.7 3.78 11.0

Source: Kirkham, 2025.

 

Table 11-3 above shows intervals that intersect the high grade are primarily encountered within the Vt unit, as would be expected. The Vt unit which represents the majority of the very high-grade populations, has 7,554 gold (3,716.8m) and 7,553 (3,716.7 m) silver assay intersections, resulting in an average grade of 9.94 g Au/t and 38.92 g Ag/t. The coefficient of variation is relatively high with 3.3 for gold and 4.0 for silver. These are reviewed once compositing

 

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and cutting is applied which will reduce the CV to reasonable values. Also, of particular interest within the Cross-cutting group are the Stock which shows 2,899 values (3,714 m) with 1.64 g/t gold and 8.11 g/t silver and HBX shows 1592 values (1,067 m) with 1.08 g/t gold and 6.94 g/t silver, respectively. The grades within the Stock and Hbx intervals display very high variability due to a small number of very high-grade outliers. These values are fairly widely distributed within the Salinas and Mita units which may positively skew the grades within the low-grade envelopes. However, as they are disseminated and to be treated within the domains, they will be cut appropriately to ensure that they reasonably represent the estimated grades.

 

Table 11-3: Statistics for weighted gold & silver assays for quaternary and cross-cutting rock types

 

Code Lith Code Metal Valid Length (meters) Max (gpt) Mean (gpt) CV
10 Qc 10 AU 787 1,271.0 5.1 0.05 2.9
AG 786 1,270.6 35 0.97 2.1
11 Qb 11 AU 144 214.7 0.06 0.01 0.4
AG 144 214.7 1 0.83 0.4
30 Cbx 30 AU 4,016 4,466.6 1,380.0 0.78 14.3
  AG 4,016 4,466.6 2,194.0 3.86 5.4
40 Gr 40 AU 419 685.1 0.246 0.01 1.5
  AG 419 685.1 2.3 0.81 0.5
50 Ad 50 AU 1,780 2,268.6 313.97 0.47 13.3
  AG 1,780 2,268.6 801.2 2.73 7.8
60 Rp 60 AU 2,899 3,714.1 46.3 0.22 2.9
  AG 2,899 3,714.1 241 2.12 2.9
70 Vt 70 AU 7,554 3,716.8 1,380.0 9.94 3.3
  AG 7,553 3,716.7 4,677.8 38.92 4.0
71 Stock 71 AU 2,383 2,214.9 148.75 1.64 3.7
  AG 2,383 2,214.9 409 8.11 2.6
72 Hbx 72 AU 1,592 1,067.4 266.09 1.08 7.9
  AG 1,591 1,067.3 969 6.94 4.7
80 RF 80 AU 5,494 6,923.0 150.7 0.28 9.2
AG 5,489 6,919.0 8,656.7 5.11 26.6
81 SZ 81 AU 36 31.5 8.4 0.27 3.0
AG 36 31.5 55.7 2.49 2.2

Source: Kirkham, 2025.

 

Table 11-4: Statistics for weighted gold & silver assays for the Salinas Group rocks

 

Code Lith Metal Valid Length (meters) Max (gpt) Mean (gpt) CV
90 Ss AU 4,200 6,269.7 15.67 0.27 2.1
AG 4,198 6,269.2 187.8 1.52 2.7
91 Svc AU 19,081 24,245.9 131.6 0.48 4.0
AG 19,032 24,189.7 1,346.9 3.41 3.9
92 Srt AU 1,215 1,522.1 16.47 0.27 2.3
AG 1,215 1,522.1 88 2.38 2.1
93 Sfx AU 1,495 2,334.3 194.7 0.34 10.8
AG 1,495 2,334.3 267.4 2.59 4.4
94 Slt AU 273 399.3 9.06 0.40 2.2
AG 273 399.3 74 1.47 4.0
95 Sct AU 242 347.7 3.57 0.19 1.9
AG 242 347.7 32 1.42 1.6
96 Scgl AU 3,189 3,481.8 157.43 0.71 3.8
AG 3,189 3,481.8 1,552.0 4.15 5.7
Total   AU 29,695 38,600.8 194.7 0.45 4.4
  AG 29,644 38,544.1 1,552.0 3.04 4.3

 

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Source: Kirkham, 2025.

 

Table 11-4 lists the statistics for the Salinas Group rocks units with the predominant unit being the Volcanic Sediments (Svc) showing mean gold and silver grades of 0.48 g/t and 3.41 g/t, respectively with relatively high variability (CV) of 4.0 and 3.9. It is apparent from logging and modeling of the Salinas that the Sinter (Ss) and the Basal Conglomerate (Scgl) illustrate consistency and continuity. In addition, the Sinter has relatively lower grades with a mean of 0.27 g/t gold while the Basal Conglomerate results show higher grades with a mean of 0.71 g/t gold as illustrated in Figure 11-2. Therefore, observations and statistical analysis supports the resultant domaining for the Salinas of the Sinter, Basal Conglomerate and the remaining sedimentary units with the Volcanic Sediments (Svc) the predominant rock type.

 

 

Figure 11-2: Box plot gold assays for the Salinas Group rocks

 

Source: Kirkham, 2025.

 

Table 11-5: Statistics for weighted gold & silver assays for the Mita Group rocks

 

Code Lith Metal Valid Length (meters) Max (gpt) Mean (gpt) CV
100 Mss AU 10,292 12,214.2 368.33 0.33 12.0
AG 10,292 12,214.2 2,405.90 2.37 10.8
101 Mat AU 5,303 6,472.7 105.647 0.45 7.1
AG 5,303 6,472.7 1,257.0 2.70 5.9
102 Mlt AU 3,387 4,703.9 62.059 0.36 5.7
AG 3,387 4,703.9 419 2.26 4.7
103 Mbt AU 22,353 24,157.8 1,380.0 0.61 10.0
AG 22,353 24,157.8 2,863.0 3.71 7.0
104 Msc AU 3,183 3,988.7 180.73 0.34 8.4
AG 3,183 3,988.7 624.6 2.20 5.5
105 Mls AU 2,750 2,981.8 163.3 0.59 6.7
AG 2,750 2,981.8 1,202.00 3.73 6.8
106 Mcv AU 21,432 28,724.3 287.13 0.32 9.9
AG 21,422 28,710.9 997.7 1.49 4.5
107 Mvo AU 2,488 2,192.1 210.3 0.53 7.5
AG 2,488 2,192.1 271 1.94 2.9
108 Mlm AU 988 852.2 45 0.37 3.4
AG 988 852.2 50.6 2.08 1.7
120 Silt AU 2 6.1 0 0.00  n/a
AG 2 6.1 0 0.00  n/a
130 PA AU 497 388.5 132.9 0.29 7.0

 

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Code Lith Metal Valid Length (meters) Max (gpt) Mean (gpt) CV
    AG 497 388.5 125 1.56 2.4
190 Mlm AU 98 73.0 14.9 0.86 2.6
AG 98 73.0 101 6.08 1.7
Total   AU 72,773 86,755.4 1,380.0 0.43 9.9
  AG 72,763 86,742.0 2,863.0 2.49 7.5

Source: Kirkham, 2025.

 

Table 11-5 lists the statistics for the Mita Group rocks units with the predominant unit being the Sandstone (Mss), Crystal Lithic Tuff (Mcv) and Lapilli Tuff (Mbt) units showing mean gold grades of 0.33 g/t, 0.61 g/t, 0.32 g/t and silver grades of 2.37 g/t, 1.49 g/t, 3.71 g/t, respectively. It is noted that the variability is very high with CV’s ranging from 4.5 to 12.0. It is again clear from logging and modeling of the Mita that the Mbt and the Mcv represent the main stratigraphic units which are distinct and significant showing consistency and continuity throughout Era Dorada.

 

Figure 11-3 shows that the Lapilli Tuff (Mbt), Conglomerate (Mvo) and Siltstone (Silt) are statistically similar, and the Upper Limestone (Mlm) is statistically different from all of the other Mita rock units. All other rock units are statistically similar as shown in Figure 11-3. Further analysis and modeling for the purpose of grouping and domaining takes these observations and conclusion into account.

  

 

Figure 11-3: Box plot gold assays for the Mita Group rocks

 

Source: Kirkham, 2025.

 

Table 11-6: Statistics for weighted gold & silver assays

 

Code Lith Metal Valid Length (meters) Max (gpt) Mean (gpt) CV
110 Tcb AU 37 49.5 0.05 0.01 1.3
AG 37 49.5 1 0.39 1.0
111 Tca AU 697 1,096.8 1.33 0.03 3.1
AG 697 1,096.8 13 0.77 1.2

Source: Kirkham, 2025.

 

Table 11-6 above shows intervals that intersect Tempisque Volcanic Complex are primarily treated as waste.

 

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11.4Geology & Domain Model

 

A three-phased modeling approach was taken to creating geology and estimation domains which included a lithostratigraphic model, detailed vein modeling, and domain modeling to estimate low-grade host rock solids within the Salinas and the Mita lithology units.

 

The lithology models were completed using the lithology codes within the database as shown in Figure 11-4.

 

 

Figure 11-4: Section view schematic of lithology for the Era Dorada Deposit

 

Source: Kirkham, 2025.

 

The models were created from first principals within LeapFrogTM and refined in MineSightTM for statistical analysis and to be used for the estimation process. Figure 11-4 illustrates the sectional interpretation of the main significant lithology units, namely the Salinas and Mita Group rock units. In addition, logging showed that within the Salinas, there appeared to be zones of gouge potentially related to fault zones termed TBX that were determined to require modeling so that they could be masked out of the domain models.

 

In addition, solid models of each of the individual veins were created and are displayed in plan in Figure 11-5 with the north veins in yellow and the south veins in blue, respectively. In preparation for the creation of the vein models, a comprehensive structural model was developed that incorporated the current drilling, underground sampling, mapping, and extensive re-logging of drill core. The models were also created from first principals using the lithostratigraphic models

 

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and the structural modeling as guides by Bluestone staff within LeapFrogTM under the supervision of the independent QP. This was done utilising the current and re- logged data, and from sectional interpretations that were subsequently wireframed based on a combination of lithology and gold grades.

 

Once completed, intersections were inspected, and all of the solids were then manually adjusted to match the drill intercepts. Once the solid models were edited and complete, they were used to code the drill hole assays and composites for subsequent statistical and geostatistical analysis. The solid zones were utilised to constrain the block model, by matching assays to those within the zones.

 

The orientation and ranges (distances) utilised for the search ellipsoids used in the estimation process were omni-directional and guided the strike and dip of the lithologic solids for the low-grade domains and by the highly constrained vein solids for the high-grade domains shown in Figure 11-5. The vein models were employed to estimate the high-grade structures on a partial block basis that are to be combined with the low-grade component to derive the whole block diluted grade for each block.

 

 

Figure 11-5: Plan view of drill holes & vein solids

 

Note: Yellow – north veins, blue – south veins.

Source: Kirkham, 2025.

 

The low-grade estimation domains were created using lithology. The methodology was to determine which lithology units could be segregated or grouped based on grade profiles and it was determined that the Salinas be modeled as Salinas, Sinter, Basal Conglomerate. Within the Mita Group, the moderately mineralised volume that envelops that North and South vein clusters are predominantly the Mbt and Mcv units.

 

Figure 11-6 and Figure 11-7 illustrate the estimation domains in the north and south, respectively, which include the veins, Salinas and Mita units.

 

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The solids were coded into the composite database in separate fields so as accurately account for the low- and high-grade components of each block along with the waste.

 

 

Figure 11-6: South area section A-A’ view of drill holes, vein solids with Salinas and Mita Units

 

Legend: Vein Solids – red; Sinter – brown polygons; Salinas – beige; Scgl conglomerate – purple; Mss – pale blue; Mat –
sapphire blue, Mbt – pale green, Mls – bright green; Mcv – dark green.

Source: Kirkham, 2025.

 

 

Figure 11-7: North area B-B’ section view of vein solids with Salinas and Mita Units

 

Legend: Vein Solids – red; Sinter – brown polygons; Salinas – beige; Scgl conglomerate – purple; Mss – pale blue; Mat
– sapphire blue, Mbt – pale green, Mls – bright green; Mcv – dark green.

Source: Kirkham, 2025.

 

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11.5Composites

 

It was determined that the 1.5 m composite lengths offered the best balance between supplying common support for samples and minimising the smoothing of grades. Figure 11-8 shows a histogram illustrating the distribution of the assay interval lengths for the complete database with 90% % of the data having interval lengths greater than 1.5 m while Figure 11-9 shows the histogram of for the assay intervals limited to within the high-grade veins where 97.5% are less than or equal to 1.5 m; 16% less than or equal to 1.0 m and 2% less than or equal to 0.5 m. To determine whether there may be selective sampling an analysis of high-grade gold samples versus assay interval lengths was performed. The scatterplot of Figure 11-10 for samples within the high-grade veins shows that the assay intervals and corresponding gold grade have the same distribution and illustrate that there is not a high-grade bias within the small intervals and sample selectivity is not occurring.

 

The 1.5 m sample length also was consistent with the distribution of sample lengths. It should be noted that although 1.5 m is the composite length, any residual composites of greater than 0.75 m in length and less than 1.5 m remained to represent a composite, while any composites residuals less than 0.75 m were combined with the composite above.

 

 

Figure 11-8: Histogram of assay interval lengths in metres

 

Source: Kirkham, 2025.

 

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Figure 11-9: Histogram of assay interval lengths within veins in metres

 

Source: Kirkham, 2025.

 

 

 

Figure 11-10: Scatterplot of assay interval lengths within veins in metres versus gold grade

 

Source: Kirkham, 2025.

 

Figure 11-11 and Figure 11-12 show histograms of the gold composite values for all composites and for those that are assigned to the high-grade veins, respectively.

 

Figure 11-13 and Figure 11-14 show histograms silver composite values for all composites and for those that are assigned to the high-grade veins, respectively. The composite data demonstrates log-normal distributions in both cases.

 

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Figure 11-11: Histogram of gold composite grades (g/t)

 

Source: Kirkham, 2025.

 

 

Figure 11-12: Histogram of gold composite grades (g/t) with vein zones

 

Source: Kirkham, 2025.

 

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Figure 11-13: Histogram of Silver Composite Grades (g/t)

 

Source: Kirkham, 2025.

 

 

 

Figure 11-14: Histogram of silver composite grades (g/t) with vein zones

 

Source: Kirkham, 2025.

 

11.5.1High-Grade Composite Analysis

 

The high-grade veins for north and south were grouped for statistical, geostatistical and estimation purposes by location and orientation in addition to relative grade profile. The results of

 

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these groupings are shown in Table 11-7 where there are 2 vein groups in the north and 6 groups in the south.

 

Table 11-7: Vein groupings for derived for statistical, geostatistical and estimation

 

Vein Domains Group Vein Ranges
VN Group 1 VN1-VN16, VN21-VN23, VN25
VN Group 2 VN17, VN18-VN20, VN24, VN26-VN30
VS Group 11 VS101 - VS103, NS121
VS Group 12 VS105-VS118
VS Group 13 VS119-VS120
VS Group 14 VS122-VS128
VS Group 15 VS132-VS138
VS Group 16 VS130-VS131, VS139

 

Source: Kirkham, 2025.

 

Statistical analysis Figure 11-15 and Figure 11-16 show the box plots and basic statistics for the grouped gold and silver composites, respectively, for the high-grade vein domains. Table 11-8 and Table 11-9 show the basic statistics for the 1.5 m gold and silver composite grades within the mineralised domains, respectively. There is a total of 6,107 composites or specifically 3,791 in the north zone and 2,316 in the south zone composites with 30 veins in the north and 36 veins in the south.

 

The weighted average gold grades for the north zone is 7.97 g/t and 7.28 g/t in the south zone with coefficients of variation (CVs) being 3.2 and 2.1, respectively. Silver grades range from 31.6 g/t in the north and 26.8 g/t in the south with CV’s being 3.4 to 3.4, respectively. CVs or variability is typically high for precious metal deposits primarily due to the nuggety nature particularly within epithermal veins; Grade limiting a cutting will further reduce the CVs.

 

The box plots and statistics show that the mean gold grade very consistent between the north and the south zones. However, the spread (i.e., SD or standard deviation) and therefore the variability (i.e., CV) are higher in the south zone. This may be due to significant outlier grades in the south which has a maximum composite value of 792.3 g Au/t which is in the very high-grade volume in VS-101 versus 276.9 g/t in VN-6 in the north. Similarly, the mean silver grades are higher in the south versus the north at 31.57 g/t and 26.77 g/t, respectfully. In addition, the silver grades have similar distribution characteristics, not only north and south but also within the individual vein groupings, with their being approximately a 4:1 ratio Ag:Au. Furthermore, variability is also significantly greater in the south which is partially due to significant outlier grades in the south where the maximum composite value is 3,540 g Ag/t in the South within VS-106 versus 1,257 g/t in the north within VN-5.

 

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Figure 11-15: Box plot of gold composites for veins

 

Source: Kirkham, 2025.

 

Table 11-8: Au composite statistics weighted by length for veins

 

Gold (g/t) Composites South North
Valid 3,791 2,316
Length 5,536 3,249.2
Minimum 0 0
Maximum 798.64 276.90
Mean 7.97 7.28
1st Quartile 0.70 0.35
Median 2.30 2.33
3rd Quartile 6.48 7.20
Standard Deviation 25.32 15.54
Variance 641.32 241.43
Coefficient of Variation 3.2 2.1

 

Source: Kirkham, 2025.

 

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Figure 11-16: Box plot of silver composites for veins

 

Source: Kirkham, 2025.

 

Table 11-9: Silver composite statistics weighted by length for veins

 

Silver (g/t) Composites South North
Valid 3,791 2,316
Length 5,536 3,249.2
Minimum 0 0
Maximum 3,539.5 1,257.0
Mean 31.57 26.77
1st Quartile 3.03 2.34
Median 8.96 6.71
3rd Quartile 25.36 21.99
Standard Deviation 108.70 72.83
Variance 11814.74 5303.76
Coefficient of Variation 3.4 2.7

 

Source: Kirkham, 2025.

 

11.5.2Low-Grade Composite Analysis

 

Figure 11-17 and Figure 11-18 show the box plots and basic statistics for the grouped (Table 11-10) gold and silver composites, respectively, for the low-grade estimation domains.

 

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Table 11-11 and Table 11-12 show the basic statistics for the 1.5 m gold and silver composite grades within the low-grade domains, respectively.

 

Table 11-10: Numeric codes for lithologies

 

CODE Litho Unit
60 Salinas (SVC)
61 Sinter (SS)
62 MAT
70 MBT
71 MCV
72 MVO
73 MAT
74 MSS
75 MLS
99 Outside

 

Source: Kirkham, 2025.

 

The low-grade envelopes show weighted average gold grades of between 0.23 and 0.55 g/t, whilst CVs between 1.6 and 5.0 show moderate to very high variability which are addressed by a conservative grade limiting and cutting strategy. It is interesting to note that the Salinas (Svc)are markedly higher grade than grade than those analysed previously which have increased from 0.19 g/t to 0.32 g/t. This may be primarily attributable updated and revised modeling of the Salinas and Sinter units which was guided by the 2021 drilling program that focussed on delineating and defining the surface resources. In addition, the Salinas Group Basal Conglomerate (Scgl) is a significantly higher-grade unit which has mean gold grade 0.55 g/t, has been defined by the updated modeling.

 

The mean Silver grades range from 1.7 to 3.4 g/t which is also lower than the 3.6 to 6.9 g/t ranges for the low-grade envelopes previously, with the CVs ranging the spectrum from low (1.2) to extreme (maximum of 39.0). As with the gold, grade limiting or cutting will further reduce the CVs. Again, it is clear that the low-grade domain composites require aggressive cutting.

 

In addition, the silver and gold grades have similar distribution characteristics, with their being an approximately a 7:1 ratio Ag:Au.

 

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Figure 11-17: Box plot of gold composites for low-grade domains

 

Source: Kirkham, 2025.

 

Table 11-11: Gold composite statistics weighted by length for low-grade domains

 

Domain Code Domain Name # Length (m) Maximum (g/t) Mean (g/t) CV
60 Svc 25,248 37,832.51 103.02 0.32 3.4
61 Ss 4,369 6,556.73 15.67 0.25 2.0
62 Scgl 3,233 4,848.43 20.79 0.55 1.6
70 Mbt 15,418 23,098.32 107.67 0.34 4.3
71 Mcv 10,487 15,718.36 52.02 0.27 4.4
72 Mvo 4,761 7,125.94 16.94 0.23 2.8
73 Mat 3,324 4,934.78 73.80 0.40 5.0
74 Mss 2,146 3,217.06 21.15 0.28 3.1
75 Mls 2,586 3,871.43 23.03 0.39 2.5
99 Outside 1,559 2,336.16 5.62 0.07 2.9

 

Source: Kirkham, 2025.

 

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Figure 11-18: Box plot of silver composites for low-grade domains

 

Source: Kirkham, 2025.

 

Table 11-12: Silver composite statistics weighted by length for low-grade domains

 

Domain Code Domain Name # Length (m) Maximum (g/t) Mean (g/t) CV
60 Svc 25,211 37,777.01 2,398.10 2.75 7.1
61 Ss 4,367 6,553.73 8,656.70 3.36 39.0
62 Scgl 3,233 4,848.43 206.9 3.36 2.0
70 Mbt 15,418 23,098.32 305.5 2.5 2.8
71 Mcv 10,486 15,717.11 251.7 1.71 3.1
72 Mvo 4,761 7,125.94 45.5 1.7 1.2
73 Mat 3,324 4,934.78 757.9 3.78 5.5
74 Mss 2,146 3,217.06 102.1 2.44 2.0
75 Mls 2,586 3,871.43 197.8 2.91 2.6
99 Outside 1,559 2,336.16 14 0.83 1.8

 

Source: Kirkham, 2025.

 

11.6Evaluation of Outlier Assay Values

 

During the estimation process, the influence of outlier composites is controlled to limit their influence and to ensure against over-estimation of metal content. The high-grade outlier thresholds were chosen by domain and are based on an analysis of the breaks in the cumulative

 

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frequency plots for each of the vein groupings and the individual low-grade domains. Figure 11-19 and Figure 11-20 show examples of the gold and silver cumulative frequency plots for all composites, respectively.

 

In the case of the gold composites, within the high-grade vein domains, values as high as 110 g/t were cut, with those as high as 500 g/t for silver cut. Table 11-13 shows the various cut thresholds for the vein groupings and Table 11-14 shows those for the low-grade domains.

 

 

 

Figure 11-19: Au cumulative frequency plot

 

Source: Kirkham, 2025.

 

 

Figure 11-20: Ag cumulative frequency plot

 

Source: Kirkham, 2025.

 

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Table 11-13: Cut grades for Au & Ag within vein domains

 

Vein Domains
Group
Domains Au Cut Threshold (g/t) Ag Cut Threshold (g/t)
VN Group 1 VN1-VN16, VN21-VN23, VN25 80 280
VN Group 2 VN17, VN18-VN20, VN24, VN26-VN30 15 40
VS Group 11 VS101 - VS103, VS121 110 180
VS Group 12 VS105-VS118 110 500
VS Group 13 VS119-VS120 10 110
VS Group 14 VS122-VS128 22 90
VS Group 15 VS132-VS138 20 95
VS Group 16 VS130-VS131, VS139 50 110

 

Source: Kirkham, 2025.

 

Table 11-14: Cut grades for Au & Ag within low-grade domains

 

Low-Grade Domain Name Domain Code Au Cut Threshold (g/t) Ag Cut Threshold (g/t)
Salinas 60 11 110
Sinter 61 6 50
SCGL 62 10 50
MBT 70 15 50
MCV 71 4.5 15
MVO 72 4.5 45.5
MAT 73 4 50
MSS 74 7 35
MLS 75 5 50
Outside 99 0.6 10

 

Source: Kirkham, 2025.

 

Table 11-15 and Table 11-16 shows the effects of cutting the outlier grades within the high-grade vein domain groupings and the low-grade Salinas and Mita units, respectively. The conclusion is that the cutting strategy is highly successful in addressing the outlier grade populations, both within the high grade veins and the lower grade Salinas and Mita units.

 

Table 11-15: Cut vs. uncut comparisons for gold and silver composites within the high-grade vein domain groupings

 

Au Maximum (g/t) Mean (g/t) CV Cut Threshold (g/t) Mean (g/t) CV Mean (g/t) CV
1 276.90 7.90 2.1 80 7.53 1.7 -5% -16%
2 66.38 3.27 1.9 15 2.87 1.4 -12% -26%
11 798.64 15.39 3.4 110 11.91 1.8 -23% -48%
12 424.15 9.95 2.2 110 9.38 1.8 -6% -19%
13 99.93 2.57 2.8 10 2.03 1.3 -21% -54%
14 95.82 3.36 2.1 22 2.99 1.4 -11% -31%
15 118.74 4.65 2.2 20 3.80 1.2 -18% -45%
16 219.40 5.09 3.4 50 3.89 1.9 -24% -43%
Total 798.64 7.70 2.9 110 6.93 2.0 -10% -32%
Ag Maximum (g/t) Mean (g/t) CV Cut Threshold (g/t) Mean (g/t) CV Mean (g/t) CV
1 1,257.0 29.68 2.6 280 26.56 1.9 -11% -28%
2 170.0 8.20 2.2 40 6.65 1.4 -19% -33%
11 1,294.5 33.52 2.7 180 26.97 1.5 -20% -44%
12 3,539.5 49.42 3.2 500 42.88 1.9 -13% -40%
13 398.2 12.14 2.4 110 10.82 1.5 -11% -40%
14 139.5 13.44 1.4 90 13.18 1.3 -2% -5%
15 343.6 16.74 1.6 95 15.55 1.3 -7% -22%
16 287.1 14.40 2.1 110 12.91 1.6 -10% -22%

 

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Au Maximum (g/t) Mean (g/t) CV Cut Threshold (g/t) Mean (g/t) CV Mean (g/t) CV
Total 3,539.5 29.75 3.3 500 26.16 2.1 -12% -37%

 

Source: Kirkham, 2025.

 

Table 11-16: Cut vs. Uncut Comparisons for Gold and Silver Composites within the Salinas and Mita Domains

 

Au Maximum (g/t) Mean (g/t) CV Cut Grade (g/t) Mean (g/t) CV Mean (%) CV (%)
60 103.02 0.324 3.4 11 0.316 2.0 -2% -41%
61 15.67 0.247 2.0 6 0.242 1.6 -2% -21%
62 20.79 0.551 1.6 10 0.546 1.5 -1% -9%
70 107.67 0.361 4.0 15 0.344 2.7 -5% -33%
71 52.02 0.260 4.5 4.5 0.223 2.4 -14% -46%
72 16.94 0.233 2.8 4.5 0.221 2.2 -5% -20%
73 73.80 0.403 5.0 4 0.306 1.8 -24% -64%
74 21.15 0.284 3.1 7 0.268 2.3 -6% -24%
75 23.03 0.388 2.5 5 0.360 1.9 -7% -24%
Total 107.67 0.327 3.7 15 0.308 2.3 -6% -38%
Ag Maximum (g/t) Mean (g/t) CV Cut Grade (g/t) Mean (g/t) CV Mean (%) CV (%)
60 2,398.1 2.75 7.1 110 2.55 2.1 -7% -70%
61 8,656.7 3.36 39.0 50 1.35 1.9 -60% -95%
62 206.9 3.36 2.0 50 3.24 1.4 -4% -32%
70 305.5 2.61 2.8 50 2.46 1.8 -6% -35%
71 251.7 1.69 3.1 15 1.43 1.4 -15% -54%
72 45.5 1.70 1.2 45.5 1.70 1.2 0% 0%
73 757.9 3.79 5.5 50 2.97 2.0 -22% -63%
74 102.1 2.44 2.0 35 2.35 1.5 -4% -21%
75 197.8 2.91 2.6 50 2.73 1.7 -6% -35%
Total 8,656.7 2.60 13.3 110 2.28 1.9 -12% -85%

 

Source: Kirkham, 2025.

 

11.7Specific Gravity Estimation

 

Table 11-17 shows the specific gravity (SG) assignment by zone using 1,308 individual measurements and standard water displacement methods. The SG assigned for the veins is determined to 2.52, which is derived from 534 measurements. There is an expanded ongoing program to increase the number and distribution of SG measurements. It is recommended that future work programs should continue to include SG measurements to expand the density distributions, particularly within the upper lithology units.

 

Table 11-17: SG zone assignments

 

Lithology Group Domain / Rock # Density (gm/mm3) Average Density (gm/mm3)
  Ss 27 2.49  
SALINAS Scgl 35 2.46  
GROUP Svc 115 2.46  
  Rp 6 2.48  
  Total 183   2.47
  Mat 48 2.54  
  Mbt 272 2.58  
MITA GROUP Mss 88 2.56  
  Mls 36 2.62  
  Mcv 102 2.59  
  Mvo 38 2.52  

 

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Lithology Group Domain / Rock # Density (gm/mm3) Average Density (gm/mm3)
  Silt 7 2.56  
  Total 591   2.57
VEIN Vt 534 2.52  
  Total 1308   2.54

 

Source: Kirkham, 2025.

 

11.8Variography

 

Experimental variograms and variogram models in the form of correlograms were generated for gold and silver grades. The definition of nugget value was derived from the downhole variograms. The correlograms for gold and silver within veins in the south and north zones are shown in Figure 11-21, Figure 11-22 and Figure 11-23 for gold and silver, respectively. These variogram models were used to estimate gold and silver grades using ordinary kriging as the interpolator used to estimate the high-grade veins.

 

 

 

Figure 11-21: Au corellogram models

 

Source: Kirkham, 2021.

 

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Figure 11-22: Ag corellogram models

 

Source: Kirkham, 2021.

 

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Figure 11-23: Ag correlogram models

 

Source: Kirkham, 2021.

 

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In addition, experimental variograms and variogram models in the form of correlograms were also generated for gold and silver grades within the low-grade domains namely, Salinas and Mita units. As above, the definition of nugget value was derived from the downhole variograms. The correlograms models for gold and silver are shown in Table 11-18 and Table 11-19, respectively. These variogram models were used to estimate gold and silver grades using ordinary kriging as the interpolator.

 

Table 11-18: Geostatistical model parameters for gold by lithology unit

 

CODE 60 61 62 70 71 72 73 74 75
Domain Name                  
Salinas Sinter MAT MBT MCV MVO MAT MSS MLS
Nugget (C0) 0.45 0.1 0.384 0.475 0.5 0.597 0.184 0.588 0.6
First Sill (C1) 0.439 0.512 0.406 0.466 0.456 0.343 0.56 0.236 0.333
Second Sill (C2) 0.111 0.388 0.21 0.059 0.044 0.059 0.256 0.176 0.067
1st Structure                  
Range along the Z' 18.1 3.6 9.7 7.2 7.8 7.9 26.9 8.9 2
Range along the X' 10.8 26.9 9.4 4.9 4.9 22.3 2.9 33.9 5.8
Range along the Y' 25.7 2.3 4.5 5.2 5.5 3.6 31.8 1.9 44.2
R1 about the Z -151 -91 -7 4 -21 15 91 1 37
R2 about the X' 35 -52 8 -37 -50 57 -47 41 -2
R3 about the Y' -4 2 -11 56 57 81 73 -42 -4
2nd Structure                  
Range along the Z' 136.6 152.6 204.4 196.5 100.8 275 76.2 12 302.3
Range along the X' 103 56.1 94.3 63.6 55 67.5 13.6 82 126.8
Range along the Y' 402.9 105.6 49.8 134.6 289 332 26.5 246 1405.4
R1 about the Z 2 24 45 2 -73 34 32 19 -15
R2 about the X' -10 56 1 24 58 171 14 41 37
R3 about the Y' -4 -23 -14 30 54 -28 33 54 41

 

Source: Kirkham, 2025.

 

Table 11-19: Geostatistical model parameters for silver by lithology unit

 

CODE 60 61 62 70 71 72 73 74 75
Domain Name                  
Salinas Sinter MAT MBT MCV MVO MAT MSS MLS
Nugget (C0) 0.4 0.231 0.3 0.425 0.167 0.462 0.35 0.279 0.274
First Sill (C1) 0.415 0.528 0.465 0.494 0.542 0.377 0.533 0.599 0.44
Second Sill (C2) 0.185 0.241 0.235 0.081 0.291 0.161 0.117 0.122 0.285
1st Structure                  
Range along the Z' 20.2 3.8 8.2 6.2 17.3 6.8 4.9 5.1 20.1
Range along the X' 4 32 3.4 9.3 8.3 17.9 30.6 37.4 7.9
Range along the Y' 8.8 2.7 33.5 4.2 3.8 43.8 19.8 2.7 1.8
R1 about the Z 1 7 -67 -34 4 23 -14 -54 -18
R2 about the X' -44 -13 87 23 -10 9 -31 -15 -1
R3 about the Y' 41 -24 20 52 -15 -22 33 -53 -20
2nd Structure                  
Range along the Z' 278.7 133.2 265.1 153 157.8 132.8 77.6 70.3 108.3
Range along the X' 45.5 10 86.3 67.6 16.8 278.3 19 115.7 13.4
Range along the Y' 70.8 89.5 73.4 208.2 27.9 71 117.9 67.3 36.7
R1 about the Z -16 8 49 42 15 7 61 -27 79
R2 about the X' 21 32 43 182 -30 35 10 90 15
R3 about the Y' 71 -8 21 -39 36 -44 -39 -5 -20

 

Source: Kirkham, 2025.

 

11.9Block Model Definition

 

The block model used for estimating the resources was defined according to origin and orientation shown in Figure 11-24 and the limits specified in Figure 11-25.

 

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Figure 11-24: Block model origin & orientation

 

Source: Kirkham, 2025.

 

Figure 11-25: Block model extents & dimensions

 

Source: Kirkham, 2025.

 

The block model employs whole blocking for ease of mine planning and is orthogonal and non-rotated, roughly reflecting the orientation of the north and the south vein sets within the deposit. The block size chosen was 5 m by 5 m by 5 m. Note that MineSight™ uses the centroid of the blocks as the origin.

 

11.10Resource Estimation Methodology

 

The estimation strategy was a two-step process that entailed estimating the high-grade vein component of each block and then the low-grade mineralised host rock component. Once completed, the final whole block grades were created by determined by way of a weighted average calculation.

 

The estimation plan for the high-grade vein component was:

 

·vein code of modelled mineralization stored in each block along with partial percentage

 

·specific gravity estimation for the vein

 

·block gold and silver grade estimation by ordinary kriging

 

·one pass estimation for each individual vein using hard boundaries

 

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A minimum of three composites and maximum of nine composites, and a maximum of three composites per hole were used to estimate block grades. Following Herco analysis, it was determined there is an appropriate amount of smoothing.

 

For the vein domains that make up the Era Dorada deposit, the search ellipsoids are omni-directional to a maximum of 100 m; however, hard boundaries were used so that the domains are tightly constrained and grade is not smeared between veins.

 

The estimation plan for the low-grade mineralised host rock component included:

 

·domain code of modelled mineralization stored in each block

 

·specific gravity estimation based on rock type code

 

·block gold and silver grade estimation by ordinary kriging

 

·one pass estimation for each domain using hard boundaries

 

A minimum of three composites and maximum of twelve composites, and a maximum of three composites per hole were informed to estimate block grades. Following Herco analysis, it was determined there is an appropriate amount of smoothing for the low-grade domains.

 

For the vein domain domains that make up the Era Dorada deposit, the search ellipsoids are omni- directional to a maximum of 100 m, and hard boundaries were used so that grade is not smeared between the units.

 

11.11Mineral Resource Classification

 

Mineral resources were estimated in conformity with generally accepted CIM “Estimation of Mineral Resource and Mineral Reserve Best Practices” Guidelines (2019). Mineral resources are not mineral reserves and do not have demonstrated economic viability. Mineral resources for the Era Dorada deposit were classified according to the CIM Definition Standards for Mineral Resources and Mineral Reserves (2014) by Garth Kirkham, P.Geo., of Kirkham Geosystems Ltd. (KGL), an Independent Qualified Person.

 

The mineral resources may be impacted by further infill and exploration drilling that may result in an increase or decrease in future resource evaluations. The mineral resources may also be affected by subsequent assessment of mining, environmental, processing, permitting, taxation, socio-economic and other factors.

 

Mineral resource categories can be based on an estimate of uncertainty within a theoretical measure of confidence. The thresholds for the uncertainty and confidence are based on rules of thumb, however they can vary from project to project depending upon the risk tolerance that the project and the company is willing to bear. Indicated resources may be estimated so the uncertainty of yearly production is approximately ±15% with 90% confidence and Measured resources may be estimated so the uncertainty of quarterly production is no greater than ±15% with 90% confidence. The results presented above indicate the reliability is around ±15% for the assumed production rate at roughly 50 m spacing.

 

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It should also be noted that the confidence limits only consider the variability of grade within the deposit. There are other aspects of deposit geology and geometry such as geological contacts or the presence of faults or offsetting structures that may impact the drill spacing.

 

The spacing distances are intended to define contiguous volumes and they should allow for some irregularities due to actual drill hole placement. The final classification volume results typically must be adjusted manually to come to a coherent classification scheme. The thresholds should be used as a guide and boundaries interpreted and defined to ensure continuity.

 

Drill hole spacing is sufficient for preliminary geostatistical analysis and evaluating spatial grade variability. The classification of resources was based primarily upon distance to the nearest composite; however, the multiple quantitative measures, as listed below, were inspected and taken into consideration.

 

The estimated blocks were classified according to:

 

·confidence in interpretation of the mineralised zones

 

·number of composites used to estimate a block

 

·number of composites allowed per drill hole

 

·distance to nearest composite used to estimate a block

 

·average distance to the composites used to estimate a block

 

Therefore, the following lists the spacing for each resource category to classify the resources assuming the current rate of metal production:

 

·Measured: Note that based on the CIM definitions, continuity must be demonstrated in the designation of measured (and indicated) resources. Therefore, measured resources were delineated from at least three drill holes spaced on a nominal 25 m pattern.

 

·Indicated: Resources in this category would be delineated from at least three drill holes spaced on a nominal 50 m pattern.

 

·Inferred: Any material not falling in the categories above and within a maximum 100 m of one hole.

 

To ensure continuity, the boundary between the indicated and inferred categories was contoured and smoothed, eliminating outliers and orphan blocks. The spacing distances are intended to define contiguous volumes and they should allow for some irregularities due to actual drill hole placement. The final classification volume results typically must be adjusted manually to come to a coherent classification scheme.

 

Mineral resources are classified under the categories of measured, indicated and inferred according to Canadian Institute of Mining, Metallurgy and Petroleum (CIM) guidelines. Mineral resource classification for gold was based primarily on drill hole spacing and on continuity of mineralization. Measured resources were defined as blocks within a distance to nearest composite of 25 m. Indicated resources were defined as those within a distance to three drill holes of less than ~50 m. Inferred resources were defined as those with an average drill hole spacing of less than ~100 m and meeting additional requirements. All resources are constrained in the following

 

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manner: primarily, by the continuous vein solids, secondarily, the low-grade envelope, and thirdly, by the Salinas group tertiary member. Blocks outside the aforementioned were estimated in a last pass to determine waste grade and volumes. Final resource classification shells were manually constructed on plan and sections.

 

The suggested classification parameters are roughly consistent with the past classification scheme. Classification in future models may differ, but principal differences should be due to changes in the amount of drilling.

 

11.12Stockpile Resources

 

Mineralised material from mining activities undertaken to date at Era Dorada, including ramp development and access, has been stockpiled on site and segregated for future processing. This material may be considered for inclusion within the initial years of mine production or within the ramp-up phase. However, this requires an accurate representation of the volumes and grades so a comprehensive sampling program was designed and implemented. The stockpile surfaces were surveyed to accurately determine volumes and the sampling program entailed excavating trenches on 20 m grid lines and 2 m sample intervals as shown in

 

.

 

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Figure 11-26.

 

 

 

Figure 11-26: Plan view of stockpile, sample locations & domain solids

 

Source: Kirkham, 2019.

 

Correlograms for gold and silver were created and employed to estimate the stockpile resources using ordinary kriging. The estimate was validated using nearest neighbour and inverse distance methods, illustrating good agreement of results.

 

Table 11-20 shows the volume and tonnage based on an unconsolidated specific gravity of 2.0 gm/mm3 along with gold and silver grades and metal content. These resources are classified as measured.

 

Table 11-20: Stockpile Resource estimate (Measured Resource)

 

Volume (BCM) Mine (t) Au (g/t) Ag (g/t) Au (oz) Ag (oz)
14,863 29,726 5.35 22.59 5,108 21,590

 

Source: Kirkham, 2019.

 

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11.13Mineral Resource Estimate

 

This estimate is based upon the reasonable prospect of eventual economic extraction based on continuity an underground mining shapes, using estimates of operating costs and price assumptions. The “reasonable prospects for eventual economic extraction” were tested using stope optimizations performed using Datamine Studio UG v.2.57TM based on reasonable prospects of eventual economic assumptions, as shown in Table 11-21.

 

Metal prices are based of long-term three-year forecast consensus financial institution estimates published by CIBC (Canadian Imperial Bank of Commerce).

 

Table 11-21: Parameters used for stope optimization and cut-off grade

 

Parameter Unit RPEEE UG Mining Method
LH MCF
Gold price US$/oz Au 2500
Project Parameters  
Process Recovery % 96.00%
Payable metal % 99.92%
TC/RC US$/oz Au 2.21
Royalty  
Royalty NSR % of NSR 1.05%
Guatemalan Gov't Royalty (Gross) % total payable metals revenue 1.00%
OPEX Estimates  
Mining  US$/t milled 100 115
Processing US$/t milled 32 32
Site Services US$/t milled 18 18
G&A US$/t milled 20 20
Total OPEX estimate US$/t milled 170 185
Cut-off Grade  
In-situ cut-off Au grade g/t 2.25 2.45

 

Source: GE21, 2025.

 

Figure 11-27 illustrates the gold block model along with the “reasonable prospects of eventual economic extraction” underground mining shapes.

 

The stope optimization results are used solely for testing the “reasonable prospects for eventual economic extraction” and do not represent an attempt to estimate mineral reserves.

 

Table 11-22 shows tonnage and grade in the Era Dorada deposit and includes all domains at a 2.25 g Au/t cut-off grade.

 

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Figure 11-27: Plan view of gold block model with reasonable prospects optimized mine shapes with existing underground ramps

 

Source: Kirkham, 2025.

 

Table 11-22: Resource estimate using 2.25 g Au/t cut-off

 

Resource Category Tonnes (kt) Au Grade (g/t) Ag Grade (g/t) Contained Gold (koz) Contained Silver (koz)
Measured          
Indicated 6,349 9.31 31.54 1,901 6,439
Measured & Indicated 6,349 9.31 31.54 1,901 6,439
Inferred 605 6.02 19.68 117 383

 

Notes:

The mineral resource statement is subject to the following:

1.Mineral Resources are reported in in accordance with S-K 1300.

2.Mineral resource estimates have been prepared by Garth Kirkham, P.Geo., a Qualified Person as defined by SK-1300.

3.The Mineral Resource estimate is reported on a 100% ownership basis.

4.Underground mineral resources are reported at a cut-off grade of 2.25 g Au/t. Cut-off grades are based on a assumed metal prices of US$ 2,500/oz gold and US$ 28/oz silver, and assumed metallurgical recovery, mining, processing, and G&A costs.

5.Mineral Resources are reported without applying mining dilution, mining losses, or process losses.

6.Resources are constrained within underground shapes based on reasonable prospects of economic extraction, in accordance with SK-1300. Reasonable prospects for economic extraction were met by applying mining shapes with a minimum mining width of 2.0 m, ensuring grade continuity above the cut-off value, and by excluding non-mineable material prior to reporting.

7.Metallurgical recoveries reported as the average over the life of mine and are assumed to be 96% Au and 85% Ag, respectively.

8.Bulk density is estimated by lithology and averages 2.47, 2.57 and 2.54 g/cm3 for the Salinas, Mita and mineralized vein domains, respectively.

9.Mineral resources are classified as Indicated, and Inferred based on geological confidence and continuity, spacing of drill holes, and data quality.

10.  Effective date of the mineral resource estimate is December 31, 2024.

11.  Tonnage, grade, and contained metal values have been rounded. Totals may not sum due to rounding.

12.  Mineral resources are not mineral reserves and do not have demonstrated economic viability.

Source: Kirkham, 2025.

 

Figure 11-28 illustrates a plan view of the 3-dimentional block model for the resources within the mineralized veins. Figure 11-29 through Figure 11-32 show sectional views of the high-grade veins for gold and silver in the north and south, respectively. Figure 11-33 through Figure 11-36 show sectional views of the total block model with the high-grade vein and low-grade host rock components resulting in the whole block grade for gold and silver in the north and south, respectively.

 

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Figure 11-28: Plan view of Au within veins along with existing ramp development

 

Source: Kirkham, 2025.

 

 

 

Figure 11-29: Section view of Au south zone veins

 

Source: Kirkham, 2025.

 

 

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Figure 11-30: Section view of Au block model north zone veins

 

Source: Kirkham, 2025.

 

 

 

Figure 11-31: Section view of Ag block model south zone veins

 

Source: Kirkham, 2025.

 

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Figure 11-32: Section view of Ag block model north zone veins

 

Source: Kirkham, 2025.

 

 

 

Figure 11-33: Section view of Au block model south

 

Source: Kirkham, 2025.

 

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Figure 11-34: Section view of Au block model north

 

Source: Kirkham, 2025.

 

 

 

Figure 11-35: Section view of Ag BLOCK MODEL NORTH

 

Source: Kirkham, 2025.

 

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Figure 11-36: Section view of Ag block model south

 

Source: Kirkham, 2025.

 

11.14Sensitivity of the Block Model to Selection Cut-off Grade

 

The mineral resources are sensitive to the selection of cut-off grade. Table 11-23 shows tonnage and grade in the Era Dorada deposit at different gold cut-off grades.

 

The reader is cautioned that these values should not be misconstrued as a mineral reserve. The reported quantities and grades are only presented as a sensitivity of the resource model to the selection of cut-off grade.

 

Table 11-23: Sensitivity analyses of tonnage along with Au & Ag grades at various Au cut-off grades

 

Resource Category Cut-off Tonnes (kt) Grade (Au g/t) Grade (Ag g/t) Contained Gold (koz) Contained Silver (koz)
Indicated 2 6,396 9.26 31.39 1,905 6,454
  2.25 6,349 9.31 31.54 1,901 6,439
  2.45 6,289 9.38 31.74 1,897 6,419
  2.5 6,269 9.40 31.81 1,895 6,410
  3 5,969 9.74 32.74 1,868 6,282
  3.5 5,552 10.22 34.11 1,825 6,089
  4 5,087 10.82 35.81 1,769 5,857
Inferred 2 623 5.91 19.45 118 389
  2.25 605 6.02 19.68 117 383
  2.45 587 6.13 19.88 116 375
  2.5 580 6.18 19.98 115 372
  3 522 6.56 20.63 110 346
  3.5 446 7.12 21.55 102 309
  4 399 7.53 22.12 96 284

 

Notes: The mineral resource statement is subject to the following:

1.All mineral resources have been estimated in accordance with Canadian Institute of Mining and Metallurgy and Petroleum (CIM) definitions, with an effective date of December 31, 2020.

2.Mineral Resources reported demonstrate reasonable prospect of eventual economic extraction; mineral resources are not mineral reserves and do not have demonstrated economic viability.

3.Cut-off grades are based on a price of US$ 2,500/oz gold, US$ 28/oz silver and a number of operating cost and recovery assumptions, plus a contingency.

4.Numbers are rounded.

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5.The mineral resources may be affected by subsequent assessment of mining, environmental, processing, permitting, taxation, socio-economic and other factors.

6.An inferred mineral resource has a lower level of confidence than that applying to an indicated mineral resource and must not be converted to a mineral reserve. It is reasonably expected that the majority of inferred mineral resources could be upgraded to indicated mineral resources with continued exploration.

Source: Kirkham, 2025.

 

11.15Resource Validation

 

A graphical validation was done on the block model. The purpose of this graphical validation is to:

 

·check the reasonableness of the estimated grades, based on the estimation plan and the near by composites.

 

·check the general drift and the local grade trends, compared to the drift and local grade trends of the composites.

 

·ensure that all blocks in the core of the deposit have been estimated.

 

·check that topography has been properly accounted for.

 

·check against partial model to determine reasonableness.

 

·check against manual approximate estimates of tonnage to determine reasonableness.

 

·inspect and explain potentially high-grade block estimates in the neighbourhood of extremely high assays.

 

A full set of cross-sections, long sections and plans were used to check the block model on the computer screen, showing the block grades and the composites. No evidence of any block being wrongly estimated was found; it appears that every block grade could be explained as a function of the surrounding composites and the estimation plan applied.

 

These validation techniques included the following:

 

·visual inspections on a section-by-section and plan-by-plan basis.

 

·the use of grade-tonnage curves.

 

·swath plots comparing kriged estimated block grades with inverse distance and nearest neighbour estimates.

 

·an inspection of histograms of distance of the first composite to the nearest block, and the average distance to blocks for all composites used, which gives a quantitative measure of confidence that blocks are adequately informed in addition to assisting in the classification of resources.

 

·validation of the block models was also performed by estimating the resources within the vein domains using partial block where the vein solids were coded as a percentage within the blocks.

 

11.16Discussion with Respect to Potential Material Risks to the Resources

 

There are no known environmental, permitting, legal, taxation, title, socio-economic, political or other relevant factors that materially affect the resources.

  

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12MINERAL RESERVE ESTIMATES

 

Mineral resources are not Mineral Reserves and have no demonstrated economic viability. This initial assessment does not support an estimate of Mineral Reserves since a pre-feasibility or Feasibility Study is required for reporting mineral reserve estimates. This report is based on potentially mineable material (mineable tonnes and/or mineable resources).

 

Mineable tonnages were derived from the resource model described in the previous section. Measured, Indicated, and Inferred resources were used to establish mineable tonnes.

 

Inferred mineral resources are considered too speculative geologically to have economic considerations applied to them that would enable them to be categorized as Mineral Reserves, and there is no certainty that all or any part of the Mineral Resources or mineable tonnes would be converted into Mineral Reserves.

 

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13MINING METHODS

 

13.1Introduction

 

The potentially mineable resources at the Era Dorada deposit will be extracted using underground mining methods, specifically a combination of mechanized longhole stoping (LH) and cut-and-fill (MCF), utilizing both paste and rock backfill. Longhole stoping is expected to account for approximately 77% of total production, while cut-and-fill will contribute the remaining 23%.

 

Mining method selection was primarily guided by geotechnical rock quality, vein geometry, and orebody continuity. Where geotechnical or geometric conditions necessitated it, mechanized cut-and-fill (MCF) was selected. Otherwise, longhole stoping was preferred due to higher productivity and lower unit mining costs relative to MCF. The proposed mine plan is designed to achieve a targeted production rate of 1,500 t/d, with a total mine life estimated at 17 years.

 

Indicated and Inferred Mineral Resources were included in mine design and schedule optimization process supporting this initial assessment. Inferred Resources are considered too geologically speculative to apply economic parameters to allow their classification as Mineral Reserves, and there is no certainty that any portion of the Inferred Resources will be upgraded to a higher resource category. The LOM plan outlined in this initial assessment is based on a resource inventory comprising approximately 78% Indicated and 22% Inferred Resources. No Mineral Resources have been classified in the Measured category.

 

13.2Deposit Characteristics

 

High-grade mineralization at the Era Dorada deposit is hosted within laterally stacked, sub-parallel narrow veins that generally strike northeast, with average azimuth ranging from 25° to 50°. Veins dips vary, including both tabular and near-vertical geometries. However, the average dip of high-grade structures is approximately 50° to 55°. The average vein thickness ranges from 2 to 10 m, with an average spacing of 8 m between parallel structures. The potentially mineable resource ranges from 50 m at the lowest levels to 300 m near the surface. The mineralized system comprises more than 50 modeled veins with variable geometry along both strike and dip.

 

Mineralization is concentrated within two main zones, the North and South. Both zones exhibit similar vein geometry and spacing. The South zone contains a greater number of veins by volume and extends to lower elevations compared to the North Zone. The combined strike length of the mineralized zone is approximately 800 m. High-grade mineralization was identified from 540 masl to 180 masl. More than 50% of the total mineralized volume occurs between 400 masl and 480 masl. In general, lower-grade mineralization envelopes the higher-grade lenses.

 

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13.3Geotechnical Analysis and Recommendations

 

13.3.1Rock Mass Characterization

 

A geotechnical investigation was carried out by Golder in 2011 and 2012 to support underground mine design, during which 16 geotechnical drill holes were logged and sampled for laboratory strength testing. Point load testing was also performed on selected core samples.

 

In 2018, JDS conducted an independent geotechnical review using the historical data as a baseline. As part of this effort, JDS performed geotechnical face mapping at 15 underground development headings and completed geotechnical logging of two additional resource drill holes. Oriented core was also collected from five drill holes located in both the North and South Zones.

 

The geotechnical assessment for this study is supported by the following dataset:

 

·RQD and core recovery data from the resource drill hole database;

 

·Detailed logging of 16 geotechnical drill holes from the 2011/2012 campaign;

 

·Over 1,500 point load tests from 43 drill holes (2011/2012);

 

·Laboratory strength testing programs from 2011 and 2018 (UCS, Brazilian tensile strength, and elastic properties);

 

·Oriented core data from five drill holes (2018);

 

·Geotechnical face mapping at 15 underground stations (2018);

 

·3D lithologic and structural models developed by SGM (2018).

 

13.3.2Geotechnical Domains and Rock Mass Properties

 

Based on the geologic structural and lithology models and the geotechnical characterization data described above, the deposit was divided by JDS (2018) into three separate geotechnical domains. Each of the domains grouped areas of similar characterized ground conditions and overall rock mass quality, which were then used to develop geotechnical design parameters. Geologic structure and lithology were identified as the dominant factors controlling rock quality domains.

 

The three geotechnical domains are shown on an E-W cross-section in Figure 13-1 and summarized below. Table 13-1 contains a summary of the key rock mass properties derived from the 2011/2012 geotechnical core logging data. The data in Table 13-1 represent the average value of all the core runs drilled within the respective domains. Local variations will occur, but the values presented are expected to be representative of the overall rock mass behavior.

 

4.Domain 1: comprises the upper lithological units of the deposit, including the upper lapilli tuff and lower volcanic sediments, both fine-grained clastic rocks characterized by closely spaced fractures. The Salinas Conglomerate, which overlies the South Zone, is also included in this domain due to its intense clay alteration. Additionally, a structurally bounded wedge of poor-quality rock, delimited by the Upper Lapilli Tuff Fault and the Ramp Fault, is part of Domain 1. This fault block has been significantly downthrown, resulting in bedding distortion and intense fracturing. Overall, Domain 1 is of poor geomechanical quality, with

 

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heavily fractured rock and relatively low intact rock strength. The average uniaxial compressive strength (UCS), derived from point load test conversions, is 71 MPa. The mean Rock Mass Rating (RMR76) is 50, classifying the rock mass as “Fair” according to Bieniawski (1976). The corresponding average Q’ value is 1.9, placing it in the “Poor” category per Barton’s classification (1974). Geotechnical mapping from five underground stations within this domain reported RMR76 values ranging from 42 to 69 (mean of 53 – “Fair”) and Q’ values from 0.8 to 7.8 (mean of 2.4 – “Poor”).

 

5.Domain 2: comprises the middle lithological units of the deposit, including andesite tuff, lower lapilli tuff, and breccias. It also includes two sandstone lenses: one located above the andesite tuff and another below the lower lapilli tuff. Compared to Domain 1, Domain 2 exhibits significantly less fracturing and higher intact rock strength. The average uniaxial compressive strength (UCS), derived from point load test (PLT) conversions, is 78 MPa. The mean Rock Mass Rating (RMR76) is 58, which corresponds to “Fair” geomechanical quality according to the Bieniawski (1976) classification system. The mean Q’ value is 4.7, also classifying the domain as “Fair” rock mass quality per Barton’s (1974) system. Geotechnical mapping at four underground stations reported RMR76 values ranging from 55 to 81, with a mean of 65—indicating “Good” rock quality. Corresponding Q’ values ranged from 3.1 to 33, with a mean value of 6.6, placing the domain within the “Fair” category under the Barton classification.

 

6.Domain 3: comprises the lower stratigraphic units of the deposit, located beneath Domain 2. These include limestone, quartz latite crystal lithic tuff, and conglomerate, as well as a sandstone lens and quartz latite unit situated between the limestone and quartz latite. The Salinas Conglomerate in the South Zone may also be included in Domain 3, where it has undergone strong silicification. This domain is characterized by good geomechanical quality, with significantly lower fracture density and higher intact rock strength. The average uniaxial compressive strength (UCS), based on point load test (PLT) conversions, is 93 MPa, with some laboratory-tested values reaching up to 233 MPa—likely due to samples with exceptionally high silica content. Domain 3 has a mean Rock Mass Rating (RMR76) of 66, corresponding to “Good” rock mass quality according to the Bieniawski (1976) classification. The mean Q’ value is 12, which also falls under the “Good” category as per Barton’s (1974) classification system. As of the current study, there are no underground exposures in Domain 3 available for geotechnical mapping.

 

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Figure 13-1: Cross-section of geotechnical domain boundaries (looking North)

 

Source: Bluestone, 2019.

 

Table 13-1: Mean rock mass properties by domain for 2011/2012 geotechnical core logging data

 

Domain Nº. No. of Runs Weath.1 (ISRM) IRS2 (ISRM) Recov. (%) RQD (%)

UCS

(MPa)3

RMR76 Q'4
1 185 W2.6 R3.4 75 29 71 50 1.9
2 625 W2.4 R3.5 89 43 78 58 4.7
3 603 W1.9 R3.8 95 67 93    
Notes:

 

1 According to ISRM (1978) rock weathering grade index, the results indicate slightly (W2) to moderately weathered (W3) rock.

2 Logged according to ISRM (1978) intact rock strength/hardness system, recorded independent of PLT results. Values indicate medium-strong (R3) to strong (R4).

3 Mean UCS values were calculated using the PLT database and a calculated correlation factor of 21, according to ISRM's (1985) suggested methods.

4 Q’ calculated from RMR76’ values using Bieniawski’s 1989 equation (Q’ = e [RMR76 – 44]/9).

Source: Bluestone, 2019.

 

The 2011/2012 geotechnical logging was done only according to the RMR76 (Bieniawski, 1976) format requiring that a conversion be made to Q’ (Barton, 1974), which is necessary for stope and ground support design. An equation developed by Barton (1979) was used to estimate Q’ values from the RMR76 values. The conversion Q’ to/from RMR using an equation is generally not preferred over collecting the necessary information for each system independently and may not be applicable for all rock masses. However, the equation was validated for the Era Dorada rock mass by comparing the Q’ and RMR76 values collected independently by JDS during the underground geotechnical mapping program (Figure 13-2). In addition, JDS geotechnically logged two recent (2018) resource drill holes and confirmed similar Q’ data for Domains 1 and 2 compared to the converted 2011/2012 data.

 

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Figure 13-2: JDS (2018) geotechnical mapping Q’ values vs. RMR76 values

 

Source: Bluestone, 2019.

 

13.3.3In-situ Stresses

 

According to JDS, in-situ stresses at the Project have not been directly measured but were estimated using regional geological data and surface topography. Based on the World Stress Map (Heidbach et al., 2016), the nearest stress data, derived from single focal mechanism earthquake events, indicate a strike-slip to the normal faulting regime, with a maximum horizontal stress (σ₁) oriented approximately 345° azimuth. These data were rated as quality ‘C’ (±25° accuracy).

 

Given the relatively shallow depth of the planned stoping areas (200–300 m below ground surface), the maximum horizontal stress magnitude is assumed to be low. No underground indicators of high horizontal stress have been observed.

 

For design purposes, the horizontal-to-vertical stress ratio (σH/σV) is conservatively assumed to range from 1.5 to 2, with σH aligned subparallel to the vein strike. The minimum horizontal stress (σH) is assumed to range from 0.5 to 1 times σV. These assumptions were used in calculating the stress parameter A for the stope design. Additional assessment of horizontal stress conditions may be warranted during operations.

 

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13.3.4Empirical Stope Design Analysis

 

Empirical stope design was conducted using stability graph methods, where the stability number (N’), calculated from rock mass quality (Q’) and empirical factors A (stress), B (structure), and C (stope dip), is plotted against the stope face hydraulic radius.

 

Maximum stope dimensions were estimated using the Potvin (2001) method, considering the expected range of rock mass conditions across the three geotechnical domains. Results were cross-validated using the Trueman & Maw6 and Q’) were evaluated. A summary of these estimates is provided in Table 13-2.

 

Table 13-2: Design rock mass quality ranges by geotechnical domain

 

Domain Nº. Lower and Upper Ranges
RQD RMR76 Q'1
1 40 to 60 45 to 55 1 to 4
2 60 to 80 55 to 65 4 to 7
3 70 to 100 60 to 70 6 to 18

Note: 1 Q’ calculated from RMR76’ values using Bieniawski’s 1989 equation (Q’ = e [RMR76 – 44]/9).

 

Source: Bluestone, 2019.

 

Stability number (N’) calculations were based on the following empirical factors:

 

·Stress factor (A): assumed as 1, reflecting relatively strong intact rock (UCS 70–100 MPa) and low horizontal stress at shallow depths (200–300 m bgs);

 

·Joint orientation factor (B): set to 0.3, as dominant discontinuities are sub-parallel to the vein orientation;

 

·Gravity factor (C): 3.8 for 45° dipping hanging walls and 2.0 for flat backs.

 

Maximum level spacing was set at 20 m. Hydraulic radii were calculated assuming a stope height of 28 m (true height across a 45° dip).

 

The geometric inputs for stope design by domain are presented below:

 

·Domain 1:

 

oCut-and-fill stopes up to 4 m high, 5 m wide, and 50 m long.

 

·Domain 2:

 

oLongitudinal stopes up to 10 m wide without cable bolting.

 

oMaximum stable length: 20 m (45° hanging wall).

 

oTransverse stopes >8 m wide require cable bolts from top cuts.

 

·Domain 3:

 

oLongitudinal stopes up to 10 m wide without cable bolting.

 

oMaximum stable length: 20 m (45° hanging wall).

 

oTransverse stopes >8 m wide require cable bolts from top cuts.

 

13.3.5Estimates of Unplanned Dilution

 

Unplanned dilution for stope hanging walls and footwalls was estimated using the Equivalent Linear Overbreak/Slough (ELOS) method (Clark, 1998), which relates the stability

 

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number (N’) to the hydraulic radius of stope faces. Unlike stability charts, the ELOS plot presents contours of expected to overbreak thickness distributed across the stope walls.

 

The method is not applicable for small hydraulic radii (<4), where dilution is primarily governed by blasting quality. Accordingly, dilution in cut-and-fill stopes within Domain 1 and relevant areas of Domains 2 and 3 was estimated based on anticipated excavation and blasting practices. For longhole stopes in Domains 2 and 3, ELOS-based dilution estimates are summarized in Table 13-3, alongside assumed values for cut-and-fill stopes. Dilution allowances for longitudinal longhole stopes were conservatively increased relative to those for transverse stopes.

 

Table 13-3: Estimates of unplanned dilution for longhole and cut-and-fill stopes by domain

 

Mining Type Stope Wall Domain Nº.
1 (Poor) A 2 (Fair) 3 (Good)
LHOS Hanging Wall NA 0.50 0.50
Footwall NA 0.40 0.30
Cut and Fill Hanging Wall 0.50 0.25 0.15
Footwall 0.50 0.25 0.15

 

Source: Bluestone, 2019.

 

13.3.6Backfill Strength Requirements

 

Stopes are planned to be backfilled with cemented paste backfill to provide lateral confinement to stope walls and to waste rock pillars between closely spaced veins, particularly in long transverse stopes. The use of cemented backfill also allows the mining of adjacent secondary stopes without leaving rib pillars.

 

Required uniaxial compressive strengths (UCS) for the paste backfill were determined analytically using a simplified wedge analysis (Mitchell, 1983), assuming a 25 m stope height and a conservative safety factor of 2.0. These values are summarized in Table 13-4. While the safety factor may be reduced to 1.5 or 1.3 during operations, a higher factor is appropriate at the initial assessment level.

 

Numerical stress modeling of backfill behavior may be considered in future stages once site-specific data is available; however, such modeling is not recommended at the current level of study.

 

Table 13-4: Required UCS for various stope widths

 

Stope Width (m) Design Backfill UCS (kPa)
2 75
5 175
10 285
15 375
20 445
25 500

 

Source: Bluestone, 2019.

 

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13.3.7Ground Support

 

Ground support requirements were assessed using the Barton & Grimstad (1994) Q-system, which incorporates rock mass quality (Q’) and the Excavation Support Ratio (ESR) to account for the function and expected lifespan of each opening. ESR values of 1.6 were applied to permanent and man-entry ore development, while a value of 3 was used for temporary, non-entry ore headings.

 

Based on the Q-system, most permanent and temporary ore development openings require only spot bolting to remain stable. However, pattern bolting with welded wire mesh is recommended in all areas where personnel will be working to control loose rock.

 

Due to concerns regarding long-term resin degradation under sustained high temperatures at Era Dorada, Swellex bolts were selected for permanent development instead of resin-anchored bolts. Groundwater pH ranges from 7.5 to 8.5, indicating minimal corrosion risk for ground support elements.

 

Recommended support strategies for each excavation type are summarized in Table 13-5 and Table 13-6.

 

Table 13-5: Ground support recommendations for ore development

 

Development Type Design

Domain 1

 

Domain 2

 

Domain 3

 

Temporary/Ore Development Back Bolts:

LHOS (4 m x 4 m); and,

Shanty Back Cut and Fill (4 m H x 4 to 6 m W).

Bolt Type Swellex Pm12 Swellex Pm12 Swellex Pm12
Bolt Diameter (mm) 27.5 27.5 27.5
Bolt Length (m) 2.1 2.1 2.1
Bolt Spacing (m) 1.2 1.5 1.5
WWM Required # 6-gauge, 10 cm opening size

Temporary/Ore Development Wall Bolts:

LHOS (4 m x 4 m); and,

Shanty Back Cut and Fill (4 m H x 4 to 6 m W).

Bolt Type Split Sets Split Sets Split Sets
Bolt Diameter (mm) 39 39 39
Bolt Length (m) 1.8 1.8 1.8
Bolt Spacing (m) 1.2 1.5 1.5
WWM Required # 6-gauge, 10 cm WWM to within 1.5 m from floor

Temporary/Ore Development 3-Way

Intersections:

LHOS (4 m x 4 m); and,

Shanty Back Cut and Fill (4 m H x 4 to 6 m W).

Bolt Type Cable Bolts Cable Bolts Cable Bolts
Bolt Diameter (mm) Single Strand Single Strand Single Strand
Bolt Length (m) 4.0 4.0 4.0
Bolt Spacing (m) 2.0 2.0 2.0

Temporary/Ore Development 4-Way

Intersections:

LHOS (4 m x 4 m); and,

Shanty Back Cut and Fill (4 m H x 4 to 6 m W).

Bolt Type Cable Bolts Cable Bolts Cable Bolts
Bolt Diameter (mm) Single Strand Single Strand Single Strand
Bolt Length (m) 5.0 5.0 5.0
Bolt Spacing (m) 2.0 2.5 2.5

Temporary/Ore Development Back Bolts:

Shanty Back Cut and Fill (4 m H x 6 to 8 m W).

Bolt Type Cable Bolts Cable Bolts Cable Bolts
Bolt Diameter (mm) Single Strand Single Strand Single Strand
Bolt Length (m) 4.0 4.0 4.0
Bolt Spacing (m) 1.2 1.5 1.5
WWM Required # 6-gauge, 10 cm # 6-gauge, 10 cm # 6-gauge, 10 cm

Temporary/Ore Development Wall Bolts:

Shanty Back Cut and Fill (4 m H x 6 to 8 m W).

Bolt Type Split Sets Split Sets Split Sets
Bolt Diameter (mm) 39 39 39
Bolt Length (m) 1.8 1.8 1.8
Bolt Spacing (m) 1.2 1.5 1.5
WWM Required # 6-gauge, 10 cm WWM to within 1.5 m from floor

Ore Development 3-Way Intersections

Shanty Back Cut and Fill (4 m H x 6 to 8 m W).

Bolt Type Cable Bolts Cable Bolts Cable Bolts
Bolt Diameter (mm) Single Strand Single Strand Single Strand

 

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Development Type Design

Domain 1

 

Domain 2

 

Domain 3

 

  Bolt Length (m) 5.0 5.0 5.0
  Bolt Spacing (m) 2.0 2.5 2.5
  Bolt Type Cable Bolts Cable Bolts Cable Bolts
Development Type Design

Domain 1

 

Domain 2

 

Domain 3

 

Ore Development 4-Way Intersections:

Shanty Back Cut and Fill (4 m H x 6 to 8 m W).

Bolt Diameter (mm) Single Strand Single Strand Single Strand
Bolt Length (m) 6.0 6.0 6.0
Bolt Spacing (m) 2.0 2.5 2.5

Estimates of Shotcrete Required:

LHOS (4 m x 4 m); and,

Shanty Back Cut and Fill (4 m H x 4 to 8 m W).

Percent Required 5.0 0.0 0.0
Thickness (cm) 5.0 0.0 0.0
Max Distance from Floor (m) 0.0 0.0 0.0

 

Source: Bluestone, 2019.

 

Table 13-6: Ground support recommendations for permanent development

 

Development Type Design Domain 1 Domain 2 Domain 3

Permanent Development

(5 m x 5 m) and

Intersection Primary

Support

Bolt Type Swellex Pm12 Swellex Pm12 Swellex Pm12
Bolt Diameter (mm) 27.5 27.5 27.5
Bolt Length (m) 2.4 2.4 2.4
Bolt Spacing (m) 1.2 1.5 1.5
WWM Required # 6-gauge, 10 cm WWM to within 1.5 m from floor

Permanent Development 3-Way Intersections

(Secondary Support)

Bolt Type Cable Bolts Cable Bolts Cable Bolts
Bolt Diameter (mm) Single Strand Single Strand Single Strand
Bolt Length (m) 4.0 4.0 4.0
Bolt Spacing (m) 1.5 1.5 1.5

Permanent Development 4-Way Intersections

(Secondary Support)

Bolt Type Cable Bolts Cable Bolts Cable Bolts
Bolt Diameter (mm) Single Strand Single Strand Single Strand
Bolt Length (m) 5.0 5.0 5.0
Bolt Spacing (m) 2.0 2.5 2.5
Estimates of Shotcrete Required Percent Required 25.0 10.0 5.0
Thickness (cm) 5.0 5.0 5.0
Max Distance from Floor (m) 0.0 0.0 0.0

 

Source: Bluestone, 2019.

 

13.4Hydrogeology Analysis and Recommendations

 

Dewatering of the Era Dorada mine has been a limiting economic factor in mine development because of high-temperature groundwater and thermal gradients. The average static groundwater-level elevation in the Project area is approximately 462 masl. Mine dewatering activities from pumping wells and two in-tunnel sump pumps (combined discharge of about 600 gallons per minute (g/m)) have lowered the static groundwater level to approximately 440 m at the mine as of 2013. The existing location of the portals, dewatering wells, and monitoring wells are shown in Figure 13-3.

 

The current conceptual model suggests that as previous dewatering efforts at Era Dorada increased, groundwater levels declined but remained within the lower Salinas tuff-volcaniclastic sequence and upper Mita Group units. The lower hydraulic heads created by pumping likely enhanced gradient- and thermal-driven recharge through faults and interconnected fractures. There is little information pertaining to the hydrogeology of the faults to determine if they provide higher permeability conduits to flow or act as barriers limiting connection between fault zones and the overall effect on long-term water quality within the area. Regional and intermediate

 

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groundwater flow systems between the Ipala Volcano and the Rio Paz shear zone likely also contribute to gradient-driven flow to Era Dorada under pumping conditions. The regional contribution of groundwater is not fully understood and needs to be further evaluated as part of the future dewatering infrastructure installation, as evidences suggests that a broad-relatively flat drawdown cone will result and could impact potential water uses and discharges to the natural environment.

 

13.4.1Evaluation of Dewatering Rates and Number of Locations

 

Historical estimates for mine dewatering have ranged from 2,000 g/m to as high as 10,000 g/m for various mine plans and have advanced as more data becomes available to better document the interconnections of the various fault systems to the regional flow system.

 

An analysis made by JDS resulted in the northern area of the mine reaching a level of 210 m within five years of mine start-up. In the southern area, the planned mine level is 270 m, which is also attained within five years. The change in planned mining depth reduced the overall dewatering demands at Era Dorada and has allowed focused dewatering strategies to be developed.

 

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Figure 13-3: Existing location of portals, dewatering wells, monitoring wells and new dewatering well locations

 

Source: Bluestone, 2017.

 

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As part of the preliminary dewatering evaluation, a simplified numerical groundwater flow model was used to simulate potential impacts associated with groundwater pumping for mine dewatering. The numerical model (MODFLOW) used a single unconfined layer with 462 m of saturated thickness and a grid of 200 rows by 200 columns (40,000 total cells) with a cell size of 25 x 25 x 500 m. The model assumed homogeneous aquifer characteristics, with a hydraulic conductivity of 0.07285 m/day and a storage coefficient of 0.028. These hydrologic parameters are based on the average values derived from testing and analyses conducted previously by MWH (2014). A general head boundary condition was used to simulate flux into and out of the model along all periphery model cells (a total of 796 general head cells). No surface recharge or evapotranspiration was simulated. This model provides a simplified tool for the calculation of dewatering rates and volumes over time and will require further development and calibration to refine dewatering estimates and injection strategies.

 

Dewatering wells were simulated at combined rates of 2,500 g/m and 3,500 g/m to evaluate the timeframe for dewatering areas around the current mine plan. The location of the dewatering wells for both scenarios is shown in Figure 13-3 (note: 2,500 g/m wells are called the “25” series, and 3,500 g/m wells are called the “35” series). While both dewatering rates, as predicted by the model, reach the desired dewatering levels, the 2,500 g/m rate does not achieve the dewatering in the desired timeframe planned for mine advancement. Therefore, a dewatering rate of 3,500 g/m was selected for a preliminary rate to be used in this PEA. The model-predicted water levels over time are shown in Figure 13-4, with key timeframes for dewatering levels based on the mine plan provided by JDS.

 

Higher dewatering rates may be required to achieve the 210 masl dewatering level within the required time frame, depending on aquifer storage and regional connection through fault zones. These details are not known at this time and require further assessment and analysis to better confirm the ultimate dewatering requirements. The 3,500 g/m dewatering rate is based on overcoming the existing storage within the aquifer and targeted dewatering along fault zones that provide connection to the regional flow system. If regional influence is significant, dewatering rates could be over twice the focused dewatering rates of 3,500 g/m and would require more wells to achieve dewatering. The increase in dewatering rates would have an effect on the need for additional injection wells and/or water treatment capacity based on the anticipated water quality. This will need to be further evaluated in the next study phase.

 

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Figure 13-4: Simulation hydrograph – south area and central area predicted dewatering at 2,500 and 3,500 g/m

 

Source: Bluestone, 2017.

 

The 3,500 g/m scenario uses 10 existing wells and 14 new dewatering wells, pumping an average of approximately 145 g/m each (see Figure 13-4). The 10 existing wells were selected because of their proximity to the mineralized zones where dewatering is required; however, the current condition of these wells is not known. Further evaluation of the existing capacity and condition is required, and either redevelopment may be required, or if conditions have deteriorated beyond recovery, new wells may be required. The new wells will likely be 12-inch diameter casing in an 18-inch diameter borehole. Telescoping perforations starting at 26 inches will likely be required for the drilling depths. It is recommended that ten piezometers be installed in small diameter boreholes (3.345-inch diameter) and equipped with vibrating wire line transducers and temperature sensors to monitor pressure changes and temperature changes in the rock as part of the dewatering program. With the new dewatering wells, these can be phased over the first two years of planned operations. Year -1 and Year 1 would require six wells installed each, and the remaining two wells could be installed in Year 2.

 

While the model shows the dewatering being reached within the desired timeframe for mining at a rate of 3,500 g/m, it does not account for the thermal effects that will be at the site, heterogeneities of the aquifer, and faults that may connect the mine site to regional groundwater flow systems, and seasonal variations in groundwater recharge. Observations from the site suggest that the system is highly compartmentalized, which may necessitate the installation of additional wells to achieve the dewatering condition. The degree of compartmentalization will need to be measured with a vibrating wire line piezometer installed in the mine area to assess pressures and temperatures. In addition, a more regional assessment of the potential effects of

 

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dewatering on the groundwater resources of the area will be required as it is expected that a broad and flat drawdown cone will result from the proposed dewatering.

 

Furthermore, potential changes in water quality over time will need to be better understood as this will impact the need for treatment to meet discharge criteria or for the suitability of the water for reinjection. The water quality assessment will need to consider the potential effects of dewatering on overall water quality and water resources in the area.

 

Finally, it is anticipated that precipitation will continue to supply approximately 400 to 600 g/m of water to the shaft, which would need to be dewatered through sumps. Overall, the water modeling suggests that the mine plan should expect to handle (through treatment or injection) a minimum water volume in the range of 4,000 g/m.

 

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Table 13-7: Wells planned and executed

 

Well Approx. Pumping Rate to Achieve
Planned Dewatering (g/m)
Dewater well Depth for
Planning (m)
Dewater Borehole
Diameter (inches)
Casing and Screen
Diameter (inches)
Easting¹ (m) Northing (m)
CBPW-2 85 Existing Existing Existing Existing Existing
CBPW-3 93 Existing Existing Existing Existing Existing
CBPW-4 141 Existing Existing Existing Existing Existing
CBPW-5 141 Existing Existing Existing Existing Existing
CBPW-6 125 Existing Existing Existing Existing Existing
CBPW-8 146 Existing Existing Existing Existing Existing
CBPW-9 141 Existing Existing Existing Existing Existing
CBPW-11 292 Existing Existing Existing Existing Existing
CBPW-14 141 Existing Existing Existing Existing Existing
CBPW-15 130 Existing Existing Existing Existing Existing
35-CO-1 180 450 18 12 212,012.5 1,587,538
35-CO-2 165 450 18 12 212,037 1,587,538
35-CO-3 125 450 18 12 212,037.3 1,587,512
35-CO-4 175 450 18 12 212,012.5 1,587,487
35-CO-5 165 450 18 12 211,962.1 1,587,312
35-SO-1 145 450 18 12 211,886.4 1,587,188
35-SO-2 125 450 18 12 211,911.6 1,587,188
35-SO-3 135 450 18 12 211,912.8 1,587,163
35-SO-4 150 450 18 12 211,886.4 1,587,138
35-SO-5 125 450 18 12 211,912.8 1,587,138
35-SO-6 150 450 18 12 211,886.8 1,587,112
35-SO-7 125 450 18 12 211,887.5 1,587,086
35-SO-8 150 450 18 12 211,862.3 1,587,062
35-SO-9 150 450 18 12 211,888.3 1,587,062
Total 3,500 6,300        

 

Note: 1 NAD1927UTMZn16N.

Source: Bluestone, 2017.

 

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13.4.2Injection Wells

 

The existing water treatment plant at Era Dorada can treat 1,500 g/m. Considering the dewatering rates are expected to be 3,500 g/m and an additional 500 g/m for mine sump water, total water treatment volumes from dewatering will be approximately 4,000 g/m in the peak dewatering periods. This volume does not consider any contact water that will be managed at the mine site. This means a total of approximately 2,500 g/m of dewatering water will need to be managed by other means than the current water treatment plant. It is recommended that 10 injection wells be installed that are capable of managing approximately 250 g/m each and will be confirmed during future injection well testing. The injection wells would be drilled to a depth of about 150 m using a 12-inch diameter casing in an 18-inch diameter borehole located on the mine property. The preliminary locations of the injection wells are shown in Figure 13-5. The injection wells are planned to be located on the south side of the mine so that groundwater withdrawals from dewatering have less impact on the local community and in order to limit pull back of water into the dewatering zone. The injection wells will likely be much shallower than the dewatering wells and will be more focused on the shallow alluvium and upper fractured bedrock. One additional injection well is planned to handle the water from the dry stack tailings facility (DSTF) runoff pond. As with the dewatering wells, injection wells can be phased in, with six planned in Year -1 and five in Year 1. Water quality, especially high iron content in the groundwater discharged to the injection wells, could cause well fouling and present a future issue with the ability to inject water efficiently. A more detailed evaluation of water quality and suitability of the water to meet injection requirements and regulatory approvals should be completed during the next study phase.

 

Further test work will be carried out in the next phase of engineering to confirm the shallower depth of injection wells will not adversely impact groundwater quality in surrounding communities and the efficiency of the injection system.

 

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Figure 13-5: Preliminary location of injection wells

 

Source: Bluestone, 2017.

 

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13.5Mining Methods

 

Two mining methods are planned for the Era Dorada deposit: sub-level longhole (LH) and cut-and-fill (MCF). The mine will be divided into mining blocks, each comprising two to three sublevels and operating independently to enhance operational flexibility and maintain production rates.

 

Sill pillars will be established between blocks to ensure safe working conditions and support high recovery rates. Multiple mining blocks will be mined concurrently, enabling the project to achieve the targeted production rate before the spiral decline reaches its full depth. Each mining block will be extracted using an overhand (bottom-up) sequence. A typical level layout is illustrated in Figure 13-6.

 

The total average mining dilution assumed for the overall deposit was 17%, based on a combination of empirical stability analysis, ELOS-based overbreak estimates, and operational assumptions related to blasting and excavation practices. This value includes both planned and unplanned dilution across all mining methods and geotechnical domains.

 

 

Figure 13-6: Perspective view of a typical mining level

 

Source: Bluestone, 2017.

 

13.5.1Longhole Mining

 

Sub-level longhole (LH) stoping is the preferred mining method at Era Dorada due to its lower operating cost and higher productivity. It is suitable for steeply dipping, continuous vein geometries in competent ground. LH stoping will be applied where geotechnical and geometric conditions allow for efficient stope design.

 

Two LH configurations will be used: longitudinal and transverse. Longitudinal stoping will be employed in the thickest and most continuous zones of the deposit. Stopes are designed to

 

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be up to 10 m long and 20 m high, with thicknesses ranging from 2 m to 50m. The minimum stope width is 2.0 m before accounting for dilution.

 

Stopes will be mined using a bottom-up (overhand) retreat sequence toward level access. Each stope is accessed via 5 x 5 m crosscuts above and below. Where stopes exceed 15 m in width, they will be divided into multiple panels with parallel access drifts and backfilled with cemented paste.

 

At Era Dorada, selected longhole stopes will be backfilled with cemented paste fill, primarily to provide structural confinement in areas where stopes are adjacent or where vein geometry requires support for subsequent mining. In transverse stopes, a primary/secondary mining sequence will be implemented, with primary stopes being backfilled to allow the safe extraction of adjacent secondary stopes without the use of rib pillars. In longitudinal stopes, structural backfill will be placed in all mined-out stopes to ensure stability during extraction. Backfill placement will occur from the top sill using paste lines, ejector trucks, or LHDs, depending on stope access and geometry. However, not all stopes will require backfill, only those where ground conditions, sequencing, or safety considerations necessitate its use.

 

The total mining dilution for longhole stopes is estimated at 17% by mass, accounting for both primary and secondary stopes. This value is based on an assumed overbreak of 0.30 m on both the hanging wall and footwall. Dilution grades were extracted from the geological block model, and a minimum stope width of 2.0 m was applied prior to dilution. These assumptions were incorporated into the mine plan and economic model to reflect anticipated operating conditions.

 

Figure 13-7 shows a typical mining sequence for LH at Era Dorada.

 

 

Figure 13-7: Longhole open stoping

 

Source: Bluestone, 2019.

 

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13.5.2Mechanized Cut-and-Fill

 

Overhand mechanized cut-and-fill (MCF) stoping is planned for areas of lower rock quality and/or where the geometry of the orebody is not conducive to longhole (LH) stoping. MCF is a highly selective underground mining method well-suited for narrow, steeply, or shallow dipping high-grade veins in weak ground conditions.

 

Mining begins at the base of the ore block and progresses upward. Each stope lift is supported temporarily with rock bolts, followed by the placement of a cemented backfill to form a competent working floor for subsequent lifts. The backfill is designed primarily for floor support rather than full structural confinement.

 

Access between successive MCF lifts is achieved via attack ramps driven at a maximum 15% gradient from the main level access. Within a typical 20-meter level interval, three MCF lifts of 4 m each are planned. The remaining 8 m to the next sublevel are mined in retreat using LH up-holes, avoiding development beneath sill pillars and enhancing miner safety. Wherever possible, internal on-vein ramping is used to reduce waste and lower costs.

 

A minimum rib pillar spacing of 4.0 m is maintained between adjacent MCF drives. In narrower veins where this spacing is not feasible, a primary/secondary mining sequence is implemented, with primary cuts backfilled using cemented structural fill to enable safe extraction of adjacent secondary cuts.

 

A schematic of a typical stope development is displayed in Figure 13-8.

 

 

Figure 13-8: Mechanized cut-and-fill

 

Source: Bluestone, 2019.

 

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13.6Mine Design

 

13.6.1Design and Optimization

 

Mine planning for the Project was conducted by GE21 using Datamine Studio UG and Mineable Shape Optimizer (MSO) software. A long section of the complete design is shown in Figure 13-9.

 

 

Figure 13-9: Mine long section

 

Source: GE21, 2025.

 

Mine design was carried out based on a gold cut-off grade (COG) calculation specific to each mining method considered within the resource model. The COG was determined using estimated gold price, metallurgical recovery, mining, processing, general and administrative (G&A) costs, and applicable royalties. The input parameters used for COG calculation are summarized (see Table 13-8).

 

Table 13-8: Cut-off grade calculation inputs

 

Parameter Unit Value
LH MCF
Gold price US$/oz Au 2,000
Project Parameters
Process Recovery % 96.00%
Payable metal % 99.92%
TC/RC US$/oz Au 2.21
Royalty
Royalty NSR % of NSR 1.05%
Guatemalan Gov't Royalty (Gross) % total payable metals revenue 1.00%
OPEX Estimates
Mining (Underground) US$/t milled 100 115
Processing US$/t milled 32 32
Site Services US$/t milled 18 18
G&A US$/t milled 20 20
Total OPEX estimate US$/t milled 170 185
Cut-off Grade
In-situ cut-off Au grade g/t 2.82 3.07

 

Source: GE21, 2025.

 

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The MSO software was used to generate optimized stope shapes based on a set of design constraints, including minimum dip angle, stope width, and gold cut-off grade (COG). Stopes designed for mechanized cut-and-fill (MCF) were located within Geotechnical Domain 1, while longhole (LH) stopes were placed within Domains 2 and 3, where geotechnical and geometric conditions allowed.

 

Following stope optimization, mine development, and access layouts were designed to ensure practical and efficient extraction sequences. Subsequently, the production schedule was optimized using Datamine's Enhanced Production Scheduler (EPS). The scheduler prioritized early access to higher-grade zones while respecting operational constraints such as maximum lateral development advance rates, plant nameplate capacity, paste backfill placement limits, and minimum required backfill cure times.

 

Mine planning was conducted using a representative stope dimension of 20 m height by 10 m width, with a 2.82 g/t Au cut-off grade for the longhole method, and 4 m high and 5 m width with a 3.07 g/t Au cut-off grade for mechanized cut-and-fill. The full set of parameters used in the selected MSO optimization trial is summarized in Table 13-9.

 

Table 13-9: Stope optimization parameters

 

Parameter Unit Value
Block Model   April 29 2025 BM export V2
Cut-off Variable ppm AuOK
Stope Orientation Plane   YZ
Framework Bearing degrees -20/ 90 Maximum change of 90
Step X m 10 / 5 (LHS/MCF)
Step Z m 20 / 4 (LHS/MCF)
Cut-Off Au g/t 2.82 / 3.07 (LHS/MCF)
Minimum Stope depth m 2
Wall Dilution m 1 / 0.3 (LHS/MCF)
Top to Bottom Max Ratio # 2.25
Max Strike Deviation degrees 45
Minimum Dip Footwall degrees 50
Minimum Dip Hanging wall degrees 130
Max Dip Change between stopes degrees 45

 

Source: GE21, 2025.

 

The stopes resulted by MSO and productive development defined the material to be input to the mining production schedule as “Run of Mine” (ROM). This material includes mineralized material classified as indicated and inferred resources and also the waste within these shapes as diluent material.

 

The sum of these materials is named as “mineable resources” for this report and resulted in 8.9 Mt @5.01 g/t Au and 17.71 g/t Ag, after applying modifying factors: for operational mining recovery was assumed a value of 95% and for mining dilution 17%.

 

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13.6.2Access

 

The Era Dorada deposit will be accessed via two main declines: one servicing the North Zone and another the South Zone. The ramps will provide haulage routes for mineralized material and waste, serve as general access, and work as air-fresh intake paths for mine ventilation.

 

Previous exploration campaigns have resulted in the development of over 2,700 m of lateral underground workings, including two portals, declines, crosscuts, and vein drifts. These workings are developed at dimensions of 4.5 m wide x 5.0 m high and are equipped with electrical power supply, ventilation infrastructure, and air and water services.

 

The existing portals were constructed into the hillsides using steel arches, corrugated steel sheeting, and shotcrete. Surface infrastructure at each portal includes supply water tanks, compressed air tanks, and an electrical power house. Although no intake fans or ducts are installed at the portals, the underground ventilation was planned via four existing 3.0 m diameter exhaust raises, each approximately 100 m in length, fitted with square concrete collars extending 1.5 m above ground.

 

These existing workings will be fully integrated into the proposed mine plan, serving as initial access, ventilation, and production levels. Additional ramps and declines will be developed at a maximum gradient of 15%, with typical dimensions of 5.0 m × 5.0 m to accommodate 30-t haul trucks and temporary 1.4 m diameter ventilation ducts. Separate ramps will be constructed to access the deeper levels of both the North and South zones.

 

As the proposed mine does not descend much farther than 400 m below the surface, and mineralization begins near the surface, no shafts were investigated as part of this conceptual study.

 

Given that the mineralization begins near the surface and extends to a depth of less than 400 m, no vertical shafts were considered in this conceptual study.

 

13.6.3Development Types

 

Spiral ramps will provide access to each production level spaced 20 m vertically apart. The spiral ramps are driven at -15% grade and 5.0 m by 5.0 m, with a maximum curvature radius of 25 m. At each operating level, the spiral ramp will run at 0% grade for 20 m to provide equipment with better visibility and turning abilities on and off the haulage ramp.

 

Each level is serviced with a footwall drive to provide ventilation, definition drilling, and crosscut development for stopes. Access drifts and footwall drives are developed 5.0 m x 5.0 m to allow truck access and reduce haul distance of LHDs. Footwall drifts are spaced a minimum of 15 m away from LH stopes to prevent stability issues as a result of production blasting.

 

Transverse LH stopes are accessed by 4.0 m x 4.0 m cross-cuts developed from footwall drives on 7.5 m spacing. Cross-cuts are used to provide a platform for LH production drills, as well as remote mucking access for blasted material.

 

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MCF zones are accessed by attack ramps from footwall drives or the haulage ramp directly. Attack ramps are driven at a maximum 15% grade and will stack vertically to access multiple production levels from a single access point, as shown in Figure 13-5. MCF drifts are driven at 4.0 m x 4.0 m to maintain structural integrity in the lower rock quality areas for which cut-and-fill is targeted.

 

Ventilation access drifts are driven on each level to ensure fresh and exhaust air raise connections to the stoping levels. The cross-cuts are approximately 4.0 m x 4.0 m.

 

Remucks are excavated on the main ramp and footwall drives to reduce the development mucking cycle time. A maximum of 150 m separates the remucks, which are typically driven 5.0 m W x 5.0 m H x 12 m l.

 

The back at the intersection of remucks and the connecting drift will require slashing to 6.4 m H to allow full extension and dumping of the LHD bucket.

 

Water collection sumps are located on every level adjacent to the exhaust raise, and after level intersection in the main ramps. Sumps have been sized at 4.0 m x 4.0 m. Additional cut-outs will be driven beside level sumps to accommodate a portable pumping skid that will collect water from the level sumps pump directly to the main dewatering sumps.

 

Electric power centers will be located outside the access drift on each level in drifts 4.0 m H x 4.0 m W. Additional power centers will be located adjacent to major power draws, such as main dewatering sumps and cooling machines.

 

Refuge station cut-outs 4.0 m x 4.0 m will be established on every level adjacent to fresh air raises. Portable refuge chambers will move between these cut-outs as needed, depending on activity within the mine. Refuge chambers will provide sufficient capacity for all persons working in the vicinity.

 

There is no plan to develop drifts dedicated entirely to diamond drilling. Any definition of diamond drilling will be carried out from footwall drives and level cross-cuts.

 

A fresh air raise of 3.0 m in diameter will be driven to connect the access drift of each level. Two exhaust raises 3.0 m in diameter will be developed at the extent of footwall drifts on each level where possible.

 

The raises are driven via a raisebore or longhole machine, depending on height. Fresh air raises will be equipped with ladders for secondary egress. The raises are sequenced in a leapfrog pattern to enable the fresh air to be carried in the direction of the ramp progression. Some return air raises will be equipped with dewatering lines and paste delivery lines as needed to supply each level of the mine.

 

In general, long-term development will incorporate a 1.0 m radius arched back, while all temporary drifts will be driven with a flat back. In areas of poor ground, it may be required to drive stope sublevels with an arched back, as their life span is generally longer than that of an MCF drift.

 

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Figure 13-10 depicts the various drift dimensions used in the Era Dorada mine plan. Figure 13-11 and Figure 13-12 depict the general arrangement of the mine plan in a long section and plan view.

 

 

Figure 13-10: Drift profiles

 

Source: Bluestone, 2019.

 

 

Figure 13-11: Mine design plan view

 

Source: GE21, 2025.

 

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Figure 13-12: Mine design long section (looking Northwest)

 

Source: GE21, 2025.

 

The geotechnical review prepared by JDS (2018) highlighted the potential difficulty and increased support requirements involved in creating large open stopes in the weaker ground zones in the Era Dorada resource. As a result of this, the mine design has been optimized to restrict LH stoping to Domain 2 and 3, and MCF stopes extract the remaining economic resource from Domain 1 (South of upper geotech domain boundary line).

 

13.7Mine Services

 

13.7.1Mine Ventilation

 

The ventilation system for the Era Dorada operation has been designed to dilute and remove dust, diesel emissions, blast fumes and provide cooling of the mine workings. The ventilation network was modelled using Ventsim software by JDS 2018 and adapted by GE21 according new production plan. The Era Dorada deposit requires additional air to pull away excess heat and control air temperatures.

 

A total of five ventilation raises are required to ventilate the mine. The Project currently has four ventilation raises, therefore only one new ventilation raise is required. As the declines are being developed, a series of ventilation drop raises will be developed concurrently. Three ventilation raises are planned to be used as exhaust raises. The remaining two raises are planned to be utilized for fresh air intake, along with the two portal ramps. The return air raises (RAR) will be required to keep air velocity on the ramp at or below 6 m/s. The fresh air raise (FAR) will also

 

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act as a means of secondary egress. The new raise will be developed using a raise climber at a 4 m by 4 m profile. Lateral ventilation drifts at 4 m by 4 m profile will be required to follow the decline and connect the ventilation circuits to the decline and level access.

 

Minimum airflow requirements were based on expected diesel emissions of the underground mining fleet required at peak mine production. Additional airflow is used underground for general cooling. The power rating of each piece of equipment was determined, and the utilization factors representing the equipment in use at any time, were applied to estimate the amount of air required. Equipment specified for site has undergone testing by either MSHA or CANMET to determine the ventilation requirements to dilute the engine emissions to a safe working level. The volume of air required for ventilating the diesel emissions is 165 m³/s. An additional 25 m³/s is used for cooling and with a 10% safety factor, the final airflow requirements for the mine were calculated at 205 m³/s.

 

Auxiliary fans will be used to ventilate the advancing development and active production levels. Fresh air will be sourced from the FAR and distributed using the auxiliary fans through ventilation ducting to the active mine areas.

 

In order to control the underground ventilation temperatures, underground mining equipment will be purchased with sealed and air conditioned cabs. When working at the active face, portable spot coolers will be used. A chiller plant located on surface will pump cold water through insulated pipes to the spot coolers used at the active faces underground to cool the air to approximately 28° C.

 

13.7.2Water Supply

 

Service water will be required mainly for drilling, dust suppression and washing of development faces. Water will be supplied from a service water tank located close to the portal and will be gravity fed to the underground work areas via 100 mm diameter pipelines. Pressure reduction valves will be installed along the decline as required. The service water tank will be refilled with cooled underground mine water or externally sourced water.

 

13.7.3Dewatering

 

Inflows into the underground workings were estimated at 25 l/s for Year 1 through 3 and 38 l/s for Year 4 onwards to the end of the mine life.

 

The mine dewatering system is designed as two standalone systems, the northern system and southern system. Both systems have been designed to accommodate a peak flow of 30 l/s, and will use a combination of 4-inch piping and 4-inch drill holes drilled between levels to transport water. A summary of the dewatering system is summarized. The location of the sumps and pumps are show in Table 13-10.

 

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Table 13-10: Underground dewatering system

 

Sump No. Section Elevation (masl) Pumps to Pump (kw) Booster (kw)
1 Northem 412 Surface via North Ramp 104 56
2 Northem 420 N Main Sump (up and around) 43  
3 Northem 360 N Main Sump 43  
4 Northem 275 360 Sump 104 56
5 Southem 440 Surface via south Ramp 104  
6 Southem 420 South Main 43  
7 Southem 330 S 420 56  
8 Southem 330 S 390 104  
9 Southem 250 S 330 104 56
10 Southem 210 S 250 104  

 

Source: GE21, 2025.

 

13.8Unit Operations

 

13.8.1Drilling

 

Development headings are planned to be driven with electro-hydraulic two-boom jumbos. Blast holes with 48 mm diameter will be drilled to a depth of 4.88 m. The advance per round is assumed to be 4.4 m. It is envisioned that one jumbo could drill between two to three rounds per shift.

 

Production drilling for the longhole stopes will be performed by longhole drills. Blast holes with 89 mm diameter will be drilled in a fan pattern from the overcut to the undercut.

 

13.8.2Blasting

 

Development rounds will be charged by an explosives and ANFO loader. Lifter holes will be loaded with bulk emulsion. Blasting is planned to be initiated by non-electric (NONEL) detonators.

 

For longhole production blasting, bulk emulsion will be used together with NONEL detonators and Pentex boosters.

 

13.8.3Ground Support

 

After mucking and scaling is complete, ground support will be installed by a mechanized scissor bolter. Typical ground support in access development is planned to consist of 2.4 m long resin rebar bolts in the back and in the walls at a spacing ranging from 1.5 x 1.5 m in moderate and poor ground. Welded wire mesh will be installed in all ground conditions. The anticipated breakdown between good, moderate, and poor is 25%, 50%, and 25%, respectively. In intersections, 3.0 m bolts will be used for deep ground support.

 

It was assumed that 25% of the development will be in poor ground conditions, which would require shotcrete. A shotcrete machine will be used to apply shotcrete at 60 mm thickness.

 

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13.8.4Mucking

 

Blasted material from development headings will be mucked with a 6.0 yd3 (10 t) LHD directly to a haul truck or to a remuck bay. Broken material from longhole stopes will be mucked by remote control LHD.

 

13.8.5Hauling

 

30 t haul trucks will drive on the decline to surface, where they will dump the material on mineralized material or waste stockpiles in close proximity to the portal.

 

Haulage profiles for all production levels were generated to calculate equipment hours for the fleet.

 

13.8.6Backfill

 

The selected mining methods require the placement of backfill for full extraction of the mineralized zones. Stopes require the use of cemented paste backfill to provide stability to exposed backfill walls when mining the adjacent stopes. The use of paste backfill will also minimize the storage requirements for process plant tailings on surface. The paste will be mixed at a paste plant and pumped through pipelines underground to the stopes. A cement content of 8% was assumed for cemented paste fill of primary stopes. Due to the high mica content of the mineralized material, 46% of the paste recipe will be crushed and screened waste rock, and 46% will be tailings from the plant. Further test work will be required to determine the optimum cement content, curing time and achievable backfill strength.

 

Underground development waste may be used for un-cemented backfill in attack ramps and remote stopes to minimize waste haulage to surface.

 

13.9Mine Equipment

 

The mobile equipment fleet to support the mining operation is summarized in Table 13-11.

 

Table 13-11: Mobile equipment fleet

 

Equipment Avg Peak Y-1 Y1 Y2 Y3 Y4 Y5 Y6 Y7 Y8 Y9 Y10 Y11 Y12 Y13 Y14 Y15 Y16 Y17
Truck (30t/14.5 m3) 3 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 -
LHD (4.5t/2.0 m3) 2 3 3 2 3 3 3 2 3 3 2 2 2 2 2 2 2 2 2 -
LHD (6.7t/3.0 m3) 2 3 3 2 3 3 3 2 3 3 2 2 2 2 2 2 2 2 2 -
LHD (10t/4.0 m3) 1 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 -
Jumbo - 1 Boom 2 3 2 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3
Jumbo - 2 Boom 2 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3
Longhole Drill - Top Hammer 1 2 1 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2
Longhole Drill - ITH 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1
Bolter 3 4 2 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4
Exploration Drill 1 1 - 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1
Small Explosives Truck 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1
Large Explosives Truck 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1
Scissor Lift 1 2 1 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2
Shotcrete Sprayer 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1
Personnel Carrier 1 2 1 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2

 

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Equipment Avg Peak Y-1 Y1 Y2 Y3 Y4 Y5 Y6 Y7 Y8 Y9 Y10 Y11 Y12 Y13 Y14 Y15 Y16 Y17
Fuel / Lube Truck 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1
Boom Truck 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1
Electrician Truck 1 1 0 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1
Grader 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1
Utility Vehicle 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2
Backhoe 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1
Telehandler 0 1 1 1 1 1 1 1 1 1 0 0 0 0 0 0 0 0 0 0
Mechanics Truck 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1
Supervisor Truck 4 4 3 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4

 

Source: GE21, 2025.

 

13.10Mine Personnel

 

The underground mine is planned to operate on two 12-hour shifts (day shift / night shift), 365 days per year with four crews on rotation. Two crews will be on-site at any time with the other crews off-site on break. Both hourly mining and maintenance personnel and salaried supervisors and technical staff will work on the same four days on, four days off rotation.

 

Hourly personnel were estimated based on development and production rates, operation productivities and maintenance requirements.

 

Underground mining personnel requirements are summarized in Table 13-12.

 

Table 13-12: Underground mine operations personnel

 

Position Avg. Quantity Hourly/Salary
Mine Management
Mine Manager 1 Salary
Mining Superintendent 1 Salary
Maintenance Superintendent 1 Salary
Technical Services Superintendent 1 Salary
Mine Foreman 1 Salary
Mine Clerk 1 Salary
Mining Operations (Production)
Shift Supervisor 4 Salary
Blasting Supervisor 4 Salary
Trainer 10 Hourly
Blaster 7 Hourly
Development Services/Shotcrete 6 Hourly
Waste Development Miner 7 Hourly
LH Prodution Miner 7 Hourly
Scooptram Operator (Large) 19 Hourly
Haul Truck Operator 18 Hourly
Bolter Operator 7 Hourly
Jackleg/Stoper Miner 8 Hourly
Grader Operator 4 Hourly
Mining Operations (Services)
Paste Plant Operators 8 Hourly
Backfill Miner 8 Hourly
Backfill Helper 8 Hourly
Mine Electrician 8 Hourly
Mine Maintenance    
Maintenance Supervisor 1 Hourly
Maintenance Planner 1 Hourly
Heavy Equipment Mechanic 16 Hourly
Mechanic Helper 4 Hourly
Welder 8 Hourly
Electric/Hydraulic Mechanic 4 Hourly

 

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Position Avg. Quantity Hourly/Salary
Mining Technical Services
Senior Mine Engineer 1 Salary
Geotechnical Engineer 1 Salary
Chief Geologist 1 Salary
Ventilation Engineer 1 Salary
Mine Surveyor 2 Salary
Surveyor Helper 2 Salary
Geologist 4 Salary
Sampler 4 Salary
Short Term Mine Planner 1 Salary
Project Engineer 1 Salary
Long-Term Mine Planner 1 Salary
Technician 2 Salary
Total Underground 194  

 

Source: Bluestone, 2017.

 

13.11Mine Production Schedule

 

An underground mine production rate of 1,500 tpd was assumed for this conceptual study, which applied indexes considered appropriate the high degree of mechanization and productivities of the selected stoping methods and available working faces and/or stopes.

 

Table 13-13 shows the mine production schedule.

 

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Table 13-13: Mine production schedule

 

Item Unit Total Y-1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17
Material Stopes* kt 8136.47 - 475.07 468.42 438.79 442.33 426.62 452.20 481.77 476.41 518.71 540.35 538.33 547.50 547.50 547.50 547.50 547.50 139.95
Des. Prod kt 763.49 - 72.43 79.08 108.71 105.17 120.88 95.30 65.73 71.09 28.79 7.15 9.17 - - - - - -
ROM kt 8899.95 - 547.50 547.50 547.50 547.50 547.50 547.50 547.50 547.50 547.50 547.50 547.50 547.50 547.50 547.50 547.50 547.50 139.95
Prod/day tpd - - 1 500.00 1 500.00 1 500.00 1 500.00 1 500.00 1 500.00 1 500.00 1 500.00 1 500.00 1 500.00 1 500.00 1 500.00 1 500.00 1 500.00 1 500.00 1 500.00 1 500.00
Des. N. Prod kt 1966.10 150.00 172.00 134.03 119.22 127.57 117.45 97.21 97.21 116.53 134.28 150.46 155.01 164.12 104.58 115.47 10.97 - -
Mass Mov Total kt 10866.17 150.00 719.50 681.53 666.72 675.07 664.95 644.71 644.71 664.03 681.78 697.96 702.51 711.62 652.08 662.97 558.47 547.50 139.95
Gold (Au) Stope* g/t 4.99 - 5.49 4.95 4.80 4.81 4.69 5.40 4.60 4.88 4.73 4.64 4.64 4.69 5.26 5.50 5.49 5.15 4.89
Des Prod g/t 5.29 - 5.87 5.95 5.46 5.20 4.14 5.31 4.95 5.42 6.68 7.97 3.96 - - - - - -
ROM g/t 5.01 - 5.54 5.09 4.93 4.88 4.57 5.38 4.65 4.95 4.83 4.68 4.63 4.69 5.26 5.50 5.49 5.15 4.89
Stope* koz 1304.66 - 83.90 74.54 67.76 68.39 64.37 78.50 71.31 74.75 78.86 80.61 80.31 82.59 92.68 96.74 96.72 90.64 21.99
Des Prod koz 129.86 - 13.66 15.14 19.09 17.59 16.09 16.28 10.46 12.38 6.19 1.83 1.17 - - - - - -
Total koz 1434.52 - 97.56 89.68 86.84 85.98 80.46 94.78 81.77 87.13 85.05 82.44 81.47 82.59 92.68 96.74 96.72 90.64 21.99
Silver (Ag) Stope* g/t 17.48 - 25.27 23.79 28.27 20.37 18.17 25.91 19.61 22.22 16.85 13.49 13.49 13.30 11.12 10.24 10.25 14.34 17.44
Des Prod g/t 20.18 - 26.53 24.21 26.41 29.19 20.55 14.07 12.98 8.60 7.26 10.56 5.73 - - - - - -
ROM g/t 17.71 - 25.43 23.85 27.90 22.06 18.70 23.85 18.81 20.45 16.35 13.46 13.36 13.30 11.12 10.24 10.25 14.34 17.44
Stope* koz 4573.07 - 385.90 358.30 398.85 289.66 249.29 376.68 303.70 340.35 281.01 234.44 233.56 234.07 195.69 180.23 180.44 252.41 78.50
Des Prod koz 495.24 - 61.77 61.55 92.31 98.71 79.88 43.10 27.43 19.65 6.72 2.43 1.69 - - - - - -
Total koz 5068.31 - 447.67 419.85 491.16 388.37 329.17 419.78 331.13 359.99 287.73 236.87 235.25 234.07 195.69 180.23 180.44 252.41 78.50
Development Lateral m 33491.24 3 306.00 3 204.40 2 448.91 2 134.00 2 259.66 2 077.70 1 713.78 1 713.78 1 990.22 2 272.36 2 560.19 2 592.19 2 609.03 1 565.42 1 043.61 -
Vertical m 3138.68 - - 48.10 87.13 117.09 110.47 97.24 97.24 180.87 229.46 242.99 295.83 401.50 467.73 401.64 361.40
Operational m 38257.66 - 2 045.60 2 752.99 3 028.88 2 873.25 3 061.83 3 438.99 3 438.99 3 078.90 2 748.18 2 446.81 2 361.98 2 192.32 2 246.31 2 043.57 499.05
Total m 74887.58 3 306.00 5 250.00 5 250.00 5 250.00 5 250.00 5 250.00 5 250.00 5 250.00 5 250.00 5 250.00 5 250.00 5 250.00 5 202.85 4 279.46 3 488.81 860.46 - -
Plant ROM processed kt 8899.95 - 547.50 547.50 547.50 547.50 547.50 547.50 547.50 547.50 547.50 547.50 547.50 547.50 547.50 547.50 547.50 547.50 139.95
Au recov.** koz 1377.14 - 93.66 86.09 83.37 82.54 77.24 90.99 78.50 83.64 81.65 79.14 78.21 79.28 88.97 92.87 92.85 87.01 21.11
Ag recov.** koz 4308.07 - 380.52 356.87 417.49 330.12 279.80 356.81 281.46 305.99 244.57 201.34 199.96 198.96 166.34 153.20 153.37 214.55 66.72
Backfill Placed kt 2847.99 - 175.20 175.20 175.20 175.20 175.20 175.20 175.20 175.20 175.20 175.20 175.20 175.20 175.20 175.20 175.20 175.20 44.79

 

Notes: *Material applying mining recovery and/or dilution.

 

Source: GE21, 2025.

 

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During mine operation two zones, North and South, each with multiple mining horizons will be in production simultaneously. The underground mine life is estimated at 17 years of production.

 

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14PROCESSING AND RECOVERY METHODS

 

The process flowsheet for the Project was based on the conclusions previously described in Section 10. Results from test programs were used to develop the corresponding process design criteria, mechanical equipment list, flowsheets and operating costs.

 

The process plant will include:

 

·Multi-staged crushing.

 

·Two-staged grinding circuit.

 

·Gravity concentration and intensive leaching (ILR).

 

·Cyanide leaching and carbon adsorption using carbon-in-pulp (CIP).

 

·Cyanide destruction, dewatering, storage of tailings dry stacking or underground deposition as paste.

 

·Carbon acid wash, elution and regeneration.

 

·Electrowinning and refining.

 

The main design criteria adopted in designing the processing circuit are listed as follows:

 

·Nominal processing rate of 1,000 tpd, which is equivalent to 0.34 Mtpa.

 

·Crusher circuit operational performance (product of availability by utilization) – OP: 65%.

 

·Grinding and extraction circuit OP: 92%.

 

·Filtration circuit OP: 92%.

 

·Grinding circuit product size: P80 of 0.053 mm.

 

14.1Description of the Process Plant

 

The selected metallurgical process flowsheet for the industrial processing of the Project comprised of the following circuits:

 

·Crushing of run of mine (ROM) ore.

 

·Crushed ore storage and reclaim.

 

·Grinding circuit with ball mills and hydrocyclones.

 

·Gravity concentration circuit, including a scalp screen, a centrifugal concentrator, and an intensive leaching reactor.

 

·Trash screening.

 

·Pre leaching thickener.

 

·Pre-oxidation in an agitated tank sparged with oxygen to oxidize the slurry prior to leaching.

 

·Cyanide Leaching in agitated leach tanks for providing 36-hour residence time to leach gold and silver into solution.

 

·Carbon in Pulp in CIP tanks to adsorb gold and silver cyanide complexes onto the pores of activated carbon.

  

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·Activated carbon acid wash, Carbon Elution and Regeneration – Acid wash of carbon to remove inorganic foulants, elution (strip) of carbon to produce a precious gold and silver rich solution, and thermal regeneration of carbon to remove organic foulants.

 

·Gold and Silver Refining by electrowinning (sludge production), filtration, drying and refining to produce gold and silver doré.

 

·Carbon safety screening.

 

·Neutralization of residual cyanide present in tailings (SO2/Air method).

 

·Final Tailings Dewatering in a thickener followed by a filter plant to reduce final tailings moisture to 18.6%, therefore adequate for dry stacking or paste backfill.

 

·Water recirculation system.

 

·Reagent storage, preparation, and dosage systems.

 

The overall process flowsheet is presented in Figure 14-1.

 

 

Figure 14-1: Overall Process Flowsheet – Era Dorada Project

 

Source: Author, 2025.

 

14.2Design Criteria

 

Key process design criteria are summarized in Table 14-1.

 

Table 14-1: Key process design criteria

 

Processing Stage Unit Nominal Value
General    
Plant Daily Throughput tpd 1000
Plant Operational Performance % 92
Overall Au Recovery % 96
Overall Ag Recovery % 85
Crushing    
Operational Performance % 65
Grinding    
Bond Ball Mill Work Index (design) kWh/t 19.9
Bond Abrasion Index g 0.24
Classification Equipment - Hydrocyclones
Final Target Product Size (P80) mm 0.053
Gravity Concentration    
Concentrator Type - Semicontinuous Batch Centrifugal
Number of Units - 1
Feed Source - Cyclone Underflow

 

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Processing Stage Unit Nominal Value
Recovery Method - Intensive Leach Reactor (ILR)
Pre-Leach Thickening    
Thickener Underflow Concentration of Solids % w/w 50
Leaching    
Pre-Oxidation Y / N Yes
Pre-Oxidation Residence Time h 2
Dissolved Oxygen Target (DO) mg/l <20
Leach Residence Time h 36
Sodium Cyanide Consumption kg/t 0.3
Lime Consumption kg/t 1.71
CIP    
CIP Residence Time h 6
Carbon Concentration mg/l 50
Carbon Loading g Au / t carbon 2,500
Carbon Processing    
Acid Wash Type - Hydrochloric Acid
Elution Operating Temperature ⁰C 140
Elution Operating Pressure kPa 350 to 500
Smelting Furnace Type - Electric Induction Furnace
Carbon Consumption Rate kg / t Carbon Stripped 30
Cyanide Destruction    
Feed Solution, CNWAD mg/l 191
Discharge Solution, CNWAD mg/l < 1.0
SO2 Consumption g / g CNWAD 4
Lime Consumption g / g CNWAD 0.8
CuSO4-5H2O Concentration mg/l 25
Tailings Management    
Disposal Type - Dry stack/Paste
Final Moisture Content % 18.6

 

Source: Bluestone, 2019.

 

14.3Process Plant Description

 

14.3.1Crushing

 

The underground Run of Mine (ROM). will be directed to an industrial crushing plant, whose product will be conveyed to a dedicated storage bin designed to a 24-hour live storage capacity.

 

14.3.2Grinding

 

The grinding circuit will process a nominal throughput of 1,000 t/day (fresh feed), for 0.053 mm (P80) ground product. A gravity concentration circuit will be installed in the grinding circuit.

 

14.3.3Gravity Concentration and Intensive Leaching

 

A fraction of the hydrocyclone nest combined underflow will be directed to the gravity concentrator scalp screen. The scalp screen will remove oversize particles prior to gravity concentration. The screen undersize will feed a semi-continuous batch gravity concentrator.

 

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The gravity concentrate will be collected in the storage cone and subsequently leached in an intensive cyanidation (leaching) reactor (ILR) circuit. The ILR pregnant solution will be pumped to the ILR pregnant solution tank, the latter located in the gold room.

 

14.3.4Pre-Leach Thickening

 

Hydrocyclone overflow will be directed to a vibrating trash screen for removal of trash material. Oversize material will discharge into a trash bin, while screen undersize will flow by gravity to pre-leach thickener. Flocculant solutions will be added to the thickener feed to enhance thickening to a nominal product of 50% w/w solids. The thickener underflow will be pumped to the pre-oxidation tank, while the thickener overflow will flow by gravity into the process water tank for recirculating in the grinding circuit.

 

14.3.5Pre-Oxidation

 

Pre-leach thickener underflow will be pumped to pre-oxidation circuit, prior to leaching. Oxygen will be sparged into the bottom of the agitated tank and slurry will be conditioned for 2 hours to oxidize Sulfide minerals.

 

Pre-oxidation will help reduce the consumption of dissolved oxygen during cyanidation, improving metallurgical recovery. It will also reduce sodium cyanide (NaCN) consumption by preventing the formation of thiocyanate and complexing some of the heavy metals such as iron. This step will also reduce reagent consumptions in the cyanide destruction circuit.

 

14.3.6Leaching

 

The leach circuit will be designed to provide 36-hour residence time. Lime slurry will be added to the first tanks at a rate of up to 1.71 kg/t to maintain protective alkalinity at a design pH of 11.0, preventing the creation of hydrogen cyanide gas (HCN). NaCN solutions will be added to the circuit at a rate of up to 0.30 kg/t, while oxygen will be sparged in from the bottom of each tank to maintain dissolved oxygen (DO) above 20 mg/l. As the slurry progresses through the circuit, gold and silver will be leached into solution.

 

Slurry from the leach circuit will then flow by gravity to the CIP circuit for carbon adsorption.

 

14.3.7Carbon in Pulp – CIP

 

Leached slurry will flow through CIP tanks for adsorbing gold-cyanide and silver-cyanide complexes onto the pores of activated carbon at an average carbon of 50 g/l to maximize adsorption.

 

As the slurry proceeds through the circuit, metal values in the solution will progressively decrease. The carbon will be transferred countercurrent to the slurry flow to maximize precious metal recovery. Regenerated carbon, with the highest adsorption potential, will be introduced into the last CIP tank, interacting with the lowest concentrations of gold and silver. Loaded carbon,

 

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with the lowest adsorption potential, will be located in the first CIP tank, interacting with the highest concentrations of gold and silver.

 

The tailings stream from the last CIP tank will flow onto a stationary safety screen to capture any carbon particles not captured in the CIP circuit. Safety screen undersize will then be pumped to the cyanide destruction circuit.

 

14.3.8Carbon Acid Wash, Elution and Regeneration (Carbon Processing)

 

The carbon processing plant will process the loaded carbon, producing gold and silver doré.

 

14.3.8.1Carbon Acid Wash

 

Loaded carbon from the CIP circuit will flow by gravity into an acid wash vessel constructed of fiber-reinforced plastic. The carbon will be treated with a circulating 3% hydrochloric acid (HCL) solution to remove calcium deposits, magnesium, sodium salts, silica, and fine iron particles. Organic foulants, such as oils and fats, are unaffected by the acid and will be removed after the elution step in the regeneration circuit using a horizontal electric kiln.

 

After the acid wash cycle, the carbon will be directed to the elution vessel using water. Under normal operation, only one acid wash and elution cycle will take place per day.

 

14.3.8.2Elution (Carbon Stripping)

 

The carbon stripping (elution) process will use the barren strip solution to strip the loaded carbon, creating a pregnant gold and silver solution which will be pumped through the electrowinning cells for precious metal recovery. The solution exiting the electrowinning cells will be circulated back to the barren solution tank for reuse.

 

During the strip cycle, solution containing approximately 1% sodium hydroxide and 0.1% NaCN, at a temperature of 140°C, will be pumped up through the strip. Solution exiting the top of the vessel will be cooled down to below its boiling point by a recovery heat exchanger. Heat from the outgoing solution will be transferred to the incoming cold barren solution prior to passing through the solution heater.

 

14.3.8.3Carbon Regeneration

 

The carbon regeneration circuit will thermally regenerate the stripped carbon, re-activating the pores and removing any organic foulants, such as oils and fats. Fresh activated carbon will be added to account for any carbon loss during the adsorption and desorption processes.

 

A horizontal electric kiln equipped with a residual heat dryer will be utilized to treat the carbon. The regenerated carbon from the kiln will flow by gravity into the carbon quench tank

  

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where it will be cooled by fresh water and/or carbon fines water, before being pumped back to the CIP circuit.

 

To compensate for carbon losses from attrition and impact, fresh carbon will be added to the carbon attrition tank and mixed with fresh water to activate the carbon pores. The fresh carbon will then drain into the carbon quench tank and combine with the regenerated carbon discharging from the kiln.

 

14.3.9Electrowinning and Refining

 

Pregnant solutions derived both from the strip circuit and ILR will be pumped to the refinery for electrowinning, therefore resulting in a gold and silver sludge.

 

Pregnant solution will be pumped through electrowinning cells, where gold and silver will plate on the stainless steel cathodes, while the barren solution will flow into the barren return tank, before being pumped back to the barren solution tank for reuse. The sludge will then be filtered, dried and refined in an electric induction furnace, producing gold and silver doré bars.

 

14.3.10Cyanide Destruction

 

The cyanide destruction circuit will consist of mechanically-agitated tanks. Cyanide will be destroyed using the SO2/Air process. Treated slurry from the circuit will then be pumped to the final tailings thickener. The cyanide destruction circuit will treat CIP tailings slurry, process spills from various contained areas, as well as process bleeding streams.

 

Oxygen will be sparged from near the bottom of the tanks, under the agitator impeller. If necessary, lime slurry will be added for maintaining the optimum pH of 8.0–8.5. Copper sulphate (CuSO4) will also be added as a catalyst, maintaining a 25 mg/l concentration in solution. A sodium metabisulphite (SMBS) solution, at a rate of up to 789 g/t, will be added into the system as the source of SO2. The system is designed to reduce the CNWAD concentration to below 1.0 mg/l.

 

14.3.11Tailing Thickening and Filtering Circuit

 

Tailings resulting from the Detox circuit at a solids concentration of 50% w/w will be pumped to a thickener where flocculant will be added.

 

The thickener underflow will be pumped to the filtering circuit for reducing the cake to a moisture content of 18.6% (dry basis). Filtering and thickening water will be recirculated within the processing plant, whereas the filtered product will be transferred to the disposal system.

 

14.4Reagent Handling, Storage and Preparation System

 

Reagents consumed within the plant will be prepared on-site and distributed via dedicated reagent handling systems. These reagents include:

 

·Sodium cyanide (NaCN).

 

·Lime.

 

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·Lead nitrate (Pb2NO3).

 

·Hydrochloric acid (HCl).

 

·Caustic soda (NaOH).

 

·Copper sulphate (CuSO4).

 

·Sodium metabisulphite (SMBS).

 

·Flocculant.

 

·Activated carbon.

 

·Antiscalant.

 

Reagents will be received and stored in appropriate facilities. Each reagent will be prepared in accordance with occupational/environmental safety standards, preventing incompatible reagents from mixing. Storage tanks will be equipped with level indicators, instrumentation, and alarms to ensure spills do not occur during normal operation. Appropriate ventilation, fire and safety protection, eyewash stations, and Material Safety Data Sheet (MSDS) stations will be located throughout the facilities.

 

The reagents will be delivered to the thickener, leach, CIP, acid wash, elution, and cyanide destruction circuits. Dosages will be controlled by flow meters and manual control valves. The capacity of the storage tanks will be designed to handle one day of production.

 

Table 14-2 summarizes the reagents used in the process plant and the respective estimated daily consumption rates. 

 

Table 14-2: Reagent consumption

 

Reagent Delivering Form Daily Usage
NaCN 1 t bags (dry) 520 kg/d
Lime 2 t bags (dry) 2.6 tpd
Pb2NO3 50 kg bags (dry) 313 kg/d
HCL 208 l drums (liquid) 592 kg/d
NaOH 50 kg bags (dry) 184 kg/d
CuSO4 50 kg bags (dry) 127 kg/d
SMBS 500 kg bags (dry) 1.35 tpd
Antiscalant 50 kg barrels 41 kg/d
Flocculant 25 kg bags (dry) 79 kg/d
Activated Carbon 50 kg bags (dry) 120 kg/d

 

Source: Bluestone, 2019.

 

14.5Utilities and Water

 

14.5.1Air Supply / Oxygen

 

An instrument and plant air system with four compressors and associated dryers, filters, and receivers will be installed in a compressor room, the latter located inside the plant building.

 

Oxygen will be used in pre-oxidation, leach, CIP and cyanide destruction circuits and will be supplied by generation systems.

 

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14.5.2Water Supply

 

Overflow water resulting from the pre-leach and tailings thickeners, as well as from the filters, will be used as process water mainly in the grinding circuit to dilute slurry to the required densities. Treated water will supply process make-up water, gland water, reagent make-up water and cooling water services in the elution circuit.

 

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15INFRASTRUCTURE

 

15.1General

 

The Project infrastructure is designed to support the operation of a 1,000 tpd underground mine and processing plant, operating on a 24 hour per day, 7 day per week basis. Support facilities have been designed to suit local conditions and topography. The main infrastructure components include the following:

 

·5 km new site access road, including an 110 m long bridge;

 

·8.2 km new 69 kV power line;

 

·On-site substation (69 kV to 13.8 kV);

 

·Water management facilities including a flood protection levee, diversion channel, ditches and collection ponds;

 

·Process plant site pad and associated buildings;

 

·Primary crusher pad;

 

·Emergency power genset;

 

·Communications system upgrade;

 

·Rehabilitation of five existing dewatering wells;

 

·Construction of eight new dewatering wells;

 

·Construction of nine new reinjection wells;

 

·Reagent warehouse and storage facilities;

 

·Truck shop (existing facility to be used in pre-production, new shop to be constructed in Operating Year 1);

 

·Fresh / Fire water tank;

 

·Process water tank;

 

·Upgrade fuel storage facility;

 

·New helipad;

 

·Upgrade Septic system for upgrade for sewage management;

 

·Solid waste disposal facility;

 

·Drystack tailings facility (DSTF);

 

·Temporary waste rock storage facility;

 

·1.0 km North and South portal connector haul road;

 

·On-site access roads for plant and facilities;

 

Additional security facilities, including site access control station.

 

15.2General Site Layout

 

The proposed site layout has been designed to support mining and plant operations while minimizing environmental and community impacts, reducing construction costs, ensuring secure access, and optimizing operational efficiency.

 

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Existing infrastructure will be used to the greatest extent possible to reduce capital costs and construction timelines.

 

Several site facilities already established at the Project will remain in use during both construction and operations. These include administrative and technical offices, modular geology and environmental units and a new assay laboratory equipped with assets from the former Marlin Mine.

 

Security infrastructure, a first aid and emergency response center, and warehouse and maintenance shops are also in place, supporting key functions such as logistics, safety, equipment servicing, and sample processing. Together, these facilities provide adequate support for mining, processing, environmental, and administrative activities across the life of the project.

 

Project overall layout is provided in Figure 15 1. The plant site and the main infrastructure facilities arrangement are presented in Figure 15 2.

 

 

 

Figure 15-1: Overall Mine Site

 

Source: Bluestone, 2019.

 

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Figure 15-2: Plant Site General Arrangement

 

Source: Bluestone, 2019.

 

15.3Site Access Road

 

Road access to the site is currently through Asunción Mita by gravel road, crossing the river Grande de Mita using a 27-tones-bridge capacity. This access is not suitable to support mining construction and operations.

 

Road access to the site is currently via Asunción Mita along a gravel road, which crosses the Grande de Mita River using a bridge with a 27-t capacity. This existing route is inadequate to support the heavy equipment loads required for mine construction and long-term operations.

 

To address this limitation, a new access route has been planned to accommodate heavy haul traffic. The proposed road will extend 5.5 km and connect directly to the Pan-American Highway (CA-1), approximately 3 km north of Asunción Mita. As part of this upgrade, a new 80-meter-long bridge will be constructed over the El Achotal River to ensure reliable and safe access to the project site during both the construction and operational phases.

 

15.4Security

 

A new access gate and guard facility will be installed at the main site entrance, including a barrier and fenced gate for preliminary screening of all incoming traffic. A site access control building and parking lot will be located nearby, where personnel access will be managed, and only approved vehicles may proceed beyond this point. The project area will be enclosed by a combination of chain-link fencing with barbed or razor wire and barbed wire livestock fencing in remote areas. Additional fencing will be installed around sensitive infrastructure such as the

 

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warehouse, refinery sections, and substations. Security will be provided by a team of uniformed and plain-clothes personnel who will monitor the main gate, conduct roving patrols, and oversee security within the refinery and general site. Security will be intensified during the construction phase, including increased staffing and coordination with contractors and local law enforcemen Mine Offices.

 

15.5Power Supply and Distribution

 

Electrical power for the project will be sourced from the Energuate Barranca Honda Substation, located approximately 1 km south of Asunción Mita. An 8.2 km long, 69 kV single-circuit overhead transmission line will be constructed. The estimated total operating load for the site, including dewatering and reinjection systems, is 14.3 MW, with a total connected load of 25.0 MW.

 

The 69 kV will be steped down power to 13.8 kV, through transformer circuits, for primary on-site power delivery to key areas including the mine, crushing plant, process plant, paste plant, cooling systems, and well fields. Existing 4.16 kV distribution lines are designed to handle 13.8 kV and will be reused where possible. However, a new 13.8 kV overhead distribution line will be built to supply power to the mine portal, mill, and related buildings. Secondary power distribution includes 4.16 kV (medium voltage) for large equipment and 480 V (low voltage) for smaller loads. Area substations will step down power accordingly, using transformers sized based on projected loads.

 

15.5.1Emergency Power

 

. The existing on-site power plant is planned to be repurposed as the emergency power facility, using diesel generators relocated from the former Marlin Mine. Generator outputs include 4.16 kV units and 600 V units, the latter of which will be stepped up to 4.16 kV via internal transformers. The standby power system is designed to supply critical loads in the event of a grid power failure, ensuring controlled shutdown of the process plant, continued operation of essential underground fans and pumps, and provision of minimum emergency power. The total estimated emergency demand is slightly under 7 MVA for both surface and underground installations.

 

15.5.2Construction Power

 

The existing standalone generators will supply the power required during the construction phase. Temporary construction generators will be used where required to provide power to remote locations, or where distribution from the existing generators is not practical.

 

15.6Process Plant

 

The proposed process plant will occupy a footprint of approximately 150 m by 70 m. It will house grinding equipment, leaching and CCD tanks, Merrill-Crowe circuit, filtration and detox systems, a gold-silver gold room, reagent preparation facilities, a dry stack tailings filtration plant,

 

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and electrical rooms. Enclosed structures will be provided for the milling circuit, Merrill-Crowe system, refinery, electrical rooms, and reagent preparation facilities. Open-air installations will include the leaching and CCD tanks, as well as the cyanide destruction area.

 

Electrical equipment buildings, including MCCs and operator control rooms for the crusher, process plant, tailings filtration, and associated areas, will be built. These units will be compliant with all local electrical and fire safety codes. Where feasible, MCCs and control gear will be pre-installed at the factory to minimize on-site work.

 

15.7Dewatering and Reinjection

 

There are currently 10 existing dewatering wells on site that are suitable for reuse. To effectively manage groundwater inflows, 14 new dewatering wells—each 450 m deep and 12 inches in diameter—will be installed. These are expected to handle a peak surface dewatering rate of approximately 795 m³/h, supplemented by 114 m³/h from underground sumps. To accommodate the increased flow, a new cooling pond will be constructed downstream of the existing south cooling pond. The site’s water treatment plant is permitted to treat 341 m³/h; excess water will be managed via reinjection through 11 newly constructed wells, each 150 m deep and capable of receiving 57 m³/h. One reinjection well will serve as contingency capacity for extreme storm events (1-in-100-year).

 

15.8Truck shop, Warehouse, Mine Dry and Administration Buildings

 

The mine truck shop and maintenance facility will be located near the mine administration building and the North Portal, providing easy access from both the mine and the plant site. The maintenance facility will include a warehouse for parts and spares. The truck shop will contain three vehicle service bays, a general shop and weld bay, and an oil change/lubrication bay. An outdoor wash bay will also be provided. An office area within the truck shop will accommodate the mine maintenance supervisor and planner.

 

A mobile equipment parts and spares warehouse will be integrated into this facility. The building will be a steel structure with metal cladding and a concrete slab-on-grade. A 10-t service crane will be installed over the service bays.

 

15.9On-Site Water Tanks

 

A new dual-purpose fresh / fire water tank will be erected with a capacity of 640,000 l. Internal risers on all non-firewater suction lines will ensure a minimum fire water reserve of 470,000 l, allowing for approximately two hours of firefighting capability.

 

A new process water tank with 170,000 l capacity will also be erected adjacent to the fresh/fire water tank to service the process plant.

 

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15.10Bulk Fuel Storage and Delivery

 

An existing diesel fuel storage facility is already installed at site for the existing power generators. It consists of two 37,500 l tanks within a concrete containment area. The existing fuel storage facility will be expanded by installing an additional 37,500 l diesel tank and increasing the containment area to accommodate the additional tank.

 

This expanded fuel storage facility will service the underground mining and site surface fleet with capacity for approximately 14 days of mobile equipment operations or two days running all critical support loads.

 

15.11Haul Roads

 

The existing North and South portal access roads will be significantly upgraded to accommodate increased traffic. Both roads will be increased to 22 m wide and the north access road will be extended to provide access to the dry stack tailing facility.

 

Various temporary construction access roads will be made or modified from existing roads for temporary construction laydown facilities, the staged DSTF construction, and for construction access, where required.

 

15.12Communications / IT

 

The existing communications tower currently provides sufficient access to stream data to and from the site and will service the site during the initial construction period. A new fiber optic cable will be installed as part of the 69 kV powerline.

 

The site communications will be distributed by fiber optic cable among the site facilities with the power transmission infrastructure. The underground mine will use a dedicated communications system. Mobile equipment and security will also use handheld radios for communications.

 

15.13First Aid / Emergency Services

 

A qualified nurse or first aid attendant will be available on site. The first aid clinic, currently under construction, will be located adjacent to the administration building and will support future operations. The facility includes a 100 m² training room for emergency response and a dedicated storage area for medical equipment and supplies.

 

An ambulance and a fire truck will be stationed in covered parking stalls near the process plant, ready for immediate deployment. All relevant buildings will be equipped with smoke, carbon monoxide, and heat detectors, as well as appropriate chemical fire extinguishers.

 

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15.14Explosives Storage and Magazines

 

The existing explosives magazine, located west of the South Portal on the southwestern side of the property, will continue to be used throughout the life of mine. The facility has adequate capacity to support approximately 50 days of operations, with storage for up to 75,000 kg of explosives (3,000 bags at 25 kg each) and 10,000 detonators. Monthly deliveries will be scheduled according to operational demand. The storage structure is built of concrete and reinforced concrete blocks and is surrounded by protective earthen and rock berms to comply with safety standards.

 

15.15Sewage Treatment

 

Sewage water will be handled by standard septic tank collection systems using natural breakdown bioreactors prior to discharge. The sanitary waste from buildings at the plant and main infrastructure site will drain to a buried septic system area below the process area. An existing bio-reactor tank is already installed and connected with buried sewer pipe for the existing site facilities buildings. An additional unit will be installed and connected to the new facilities that are being added for the project.

 

Sewage will be treated, separated, and the liquid discharged. Water for septic operation and wash use will be made up from the raw water supply.

 

15.16Surface Water Management

 

Surface water infrastructure at the site is designed to separate and manage “contact” and “non-contact” water, minimizing the potential for contamination during mine operations. Contact water—originating from areas such as the process plant and DSTF where filtered tailings are handled—is either reused in the plant or treated at the water treatment plant (WTP). Non-contact runoff is directed to designated discharge points with sediment control capacity. Stormwater is managed through 13 lined channels routed to seven ponding areas, which include both contact and non-contact ponds. The system incorporates reinforced channels, culverts, bridges, and energy dissipation structures to prevent erosion and manage flow during storm events. Contact water ponds are sized for 100-year, 24-hour storm retention and were further evaluated using a site-wide water balance based on climate data from 1970 to 2017.

 

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Figure 15-3: Storm Water Management Infrastructure Surrounding the DSTF

 

Source: Bluestone, 2019.

 

15.17Fresh Water Supply

 

Mine water will be the primary source of fresh water for the process plant, dust control, and for treated water use for the site personnel facilities. The mine water will be treated in the existing water treatment plant prior to use as the site fresh water supply. On-site office facilities, and staff/contractor dining facilities will use the treated water for washing, laundry, and bathing. Drinking water and cooking water will be provided by purchasing potable bottled water from a local vendor.

 

15.18Water Treatment Infrastructure

 

The existing water treatment plant (WTP) at Era Dorada is permitted to treat and discharge up approximately 341 m³/h. The plant is designed to remove arsenic via co-precipitation with ferric salts, using a treatment sequence that includes chemical oxidation, pH adjustment, ferric dosing, and solids separation. It also receives up to ≈ 61 m³/h of process water bleed from the tailings thickener overflow. The facility has modular capacity, allowing for future expansion if permit limits are increased.

 

To address potential mercury and copper concentrations in effluents, proposed process modifications include the use of a sulfur-based reagent effective in removing divalent heavy metals. The treated and cooled mine water is also used as process and utility water and will be stored in a dual-purpose raw/fire water tank located near the process plant. As gravity pressure is insufficient due to elevation differences, pumps will be installed to ensure adequate pressure

 

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throughout the distribution system. The plant also plays a key role in thermal regulation by cooling mine water via a series of ponds prior to reuse or discharge.

 

The Project incorporates a water management strategy that allows for the reuse of treated water from the water treatment plant (WTP), in addition to groundwater sources. Treated water is reused in the process plant and for operational services, while excess volumes are managed according to environmental discharge and reinjection plans.

 

15.19Tailings Management Facility

 

15.19.1Drystack Tailings Facility

 

The Project will utilize a drystack tailings facility (DSTF) for tailings management, incorporating filtered tailings transport and placement to minimize environmental impacts and enhance long-term stability. The DSTF will accommodate approximately 3 Mt of tailings over the mine life, corresponding to a required volume of 1.9 million cubic meters, based on a weighted average dry density of 1.59 t/m³. The facility is configured as a centerline-raised embankment, starting with a rockfill starter dam and built progressively with compacted filtered tailings.

 

Tailings placement follows a seasonal deposition strategy: during the wet season, material will be deposited loosely in the western section, shaped to direct runoff into decant structures; in the dry season, tailings will be compacted in horizontal lifts in the eastern section to provide structural support and storage capacity. Transport to the DSTF will be carried out by a mine fleet of haul trucks. Initial capital works include the construction of the impoundment area, underdrain systems, geotextile lining, and reclaim ponds, as well as installation of the mechanical and electrical systems required to recirculate water back to the process plantt. The general layout of the DSTF is illustrated in Figure 15-4.

 

 

 

Figure 15-4: DSTF Seasonal Material Placement Plan

 

Source: Bluestone, 2019.

 

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15.19.2Subsurface Preparation

 

A series of geotechnical investigations were conducted to support the design of the DSTF. These included test pits, boreholes, Standard Penetration Tests (SPT), and in-situ permeability testing. The 2018 campaign by Stantec, complemented by prior work from Golder (2012–2013), provided a comprehensive understanding of the foundation conditions at the proposed DSTF site. The subsurface profile generally consists of colluvial and alluvial soils over residual materials derived from sedimentary and volcanic rocks. Bedrock was not encountered within the depth range investigated (up to 30.3 m).

 

This site characterization allowed for the development of foundation design criteria, material suitability assessments, and embankment performance evaluations under both static and seismic loading. The data were used to confirm the feasibility of the selected location and inform the layout and sequencing of construction activities.

 

 

 

Figure 15-5: DSTF Geotechnical Site Investigation Plan

 

Source: Bluestone, 2019.

 

15.19.3Seepage Collection System

 

To ensure long-term performance and reduce pore pressure within the tailings mass, a dedicated seepage collection system has been incorporated into the DSTF design. The system includes a foundation-level underdrain network and perforated vertical decant towers designed to intercept percolated water and surface runoff. Water collected from both systems will flow to a low-flow reclaim pond, with overflow capacity directed to a stormwater pond for exceptional rainfall events.

 

Each pond is equipped with a sump and pump system to return water to the process plant for reuse. The construction of drainage infrastructure will be staged in coordination with tailings deposition, ensuring operational readiness as the facility expands. This approach supports the

 

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project’s zero-discharge water management strategy and enhances geotechnical stability over time.

 

 

 

Figure 15-6: DSTF Underdrain Plan Showing Starter Dam

 

Source: Bluestone, 2019.

 

15.20Waste Rock Facility

 

The planned Waste Rock Facility (WRF) is located southwest of the DSTF, near the south portal, and is designed to temporarily store up to 120,000 m³ of waste rock generated during mining. Bluestone intends to use most of this material in the underground backfill program, resulting in minimal long-term accumulation. While geochemical testing is ongoing to confirm the potential for acid generation (PAG), current assumptions consider low risk due to limited exposure time. Field humidity cell tests on representative samples are recommended prior to detailed engineering.

 

Although no targeted site investigation was conducted beneath the WRF footprint, general test pits in the area indicate a subsurface profile of sands and gravels, considered typical and adequate for feasibility-level design. Observations of existing waste rock at the south portal confirm that materials are hard and durable. Further geotechnical and geochemical studies will be required in subsequent design phases to validate long-term stability and environmental compliance.

 

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Figure 15-7: WRF General Configuration Plan

 

Source: Bluestone, 2019.

 

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16MARKET STUDIES

 

16.1Gold Market

 

The global gold market operates as a well-established and highly liquid system, characterized by a diversified foundation of supply and demand. From a macroeconomic perspective, gold consistently exhibits countercyclical behavior, having historically served as a store of value under conditions of elevated financial stress, inflation volatility, and geopolitical instability. Its low to negative correlation with traditional asset classes—such as sovereign bonds and equities—significantly enhances its utility as a portfolio diversifier (Figure 16-1).

 

 

 

Figure 16-1: Gold price behavior since 2000

 

Source: World Bank Group, 2025.

 

16.1.1Gold Price

 

Mineral Resources have been modelled at a gold price of US$ 2,000/troy oz. Project economics have also been assessed at a base case gold price of US$ 2,389/troy oz based on the long-term consensus forecast from over 20 investment banks. Project economics at a range of gold prices are evaluated as part of project sensitivity analysis in Section 22.

 

16.2Silver Market

 

Relative to global markets such gold, the global silver market is less significant in value. According to data published by the Silver Institute, it reached 680.5 million ounces (Moz) in 2024 and is projected to exceed 700 Moz in 2025 (Figure 16-2).

 

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Figure 16-2: Silver price behavior since 2000

 

Source: World Bank Group, 2025.

 

16.2.1Silver Price

 

Although silver tonnage has not been explicitly modeled within the mineral resource estimate, project economics were assessed using a silver price of US$ 28.44/troy oz, based on the long-term consensus forecast from over 20 investment banks.

 

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17ENVIRONMENTAL STUDIES, PERMITTING, AND PLANS, NEGOTIATIONS, OR AGREEMENTS WITH LOCAL INDIVIDUALS OR GROUPS

 

17.1Introduction

 

Aura’s review shows that the Project has all necessary permits to proceed with the development of the underground mine and construction of the process facilities, subjected to the future operation to adhere to the conditions of the exiting permits. The review shows that the Project has been operating so far under high levels of environmental and CSR standards and has been maintaining a comprehensive Permits Register, which shows that all applicable permit commitments have been fulfilled over time.

 

17.2Environmental Impact Assessment and Permitting

 

There are various environmental studies and ongoing monitoring activities which have been performed at the Era Dorada site since the start of the project. An Environmental Impact Assessment (EIA) was submitted and approved by the Ministerio de Ambiente y Recursos Naturales (MARN) in 2007, however some components of the mine design have changed since that time and specific permit amendments are required. Additionally, it is important to highlight that new baseline environmental and social studies are required for the power line.

 

The approved EIA from 2007 included basic Environmental Management Plan (EMP), Social Management Plan (SMP) and Conceptual Mine Closure Plan, which have been reviewed and updated during the Feasibility Study (Blustone, 2019) to account for current international good-practices and the updated project design. Over the next project phase, those plans will be updated to reflect optimization and further development.

 

17.2.1EIA Areas of Influence

 

In 2007, the approved EIA was prepared based on three specific areas, as shown in Figure 17-1. The approved EIA has defined the direct area of influence within an irregular polygon 235,452 m2 in area. This includes the underground mine, the processing plant and its surrounding service buildings. The indirect area of influence includes exploration areas in addition to the direct area of influence, covering a total area of 7,050,000 m2. An external area of influence is also considered; this includes the surrounding areas and the following seven communities (total area of 5,500 ha):

 

·Caserlo La Lima;

 

·Trapiche Vargas;

 

·El Cerron;

 

·El Tule;

 

·Las Animas;

 

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·San Rafael Cerro Blanco; and.

 

·Municipality of Asunción Mita.

 

 

 

Figure 17-1: EIA Areas of Influence

 

Source: Bluestone, 2019

 

17.2.2Permitting

 

As already mentioned, since the design has been updated and optimized, an amendment of the 2007 EIA and specific permits will be required for approval to be aligned with the updated project design.

 

The power line is not covered by any previous studies or permits, therefore requiring new baseline studies, EIA, and permit applications to be submitted to MARN for approval, with input from the following Guatemalan authorities: Ministerio de Energía y Mineria (MEM), Consejo Nacional de Areas Protegidas (CONAP), Instituto Nacional de Bosques (INAB), Ministerio e Salud y Asistencia social (Ministry of Health & Social Assistance), and the local municipality of Asunción Mita. The anticipated duration for completion of baseline studies, submittal/approval of EIA, and issue of permits is 8-10 months.

 

The

 

Table 17-1 provides a summary of main permit amendments and new permits required, while Source: Bluestone, 2019.

 

Table 17-2 summarizes the ongoing permits and current status of each.

 

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Table 17-1: Main permit amendments & new permit required

 

Project Component Action Required
Water Management (Injection of Mine Water// Increase of water discharge flow) EIA Amendment
DSTF Optimization (Changes from 2007 EIA Design) EIA Amendment
Change of project footprint or approved design (Non authorized activities) EIA Amendment
New Power Line New EIA and Permit
Export Permit New Permit

 

Source: Bluestone, 2019.

 

Table 17-2: Current permits

 

License / Permit Resolution / Date of Issue Expiration Date
Mining Resolution No.1942 MEM 2032
Tracking and Surveillance Licence Category A 2613-2007/ECM/LP 2028
EIA approval 2613-2007/ECM/LP MARN The duration of the Project life
Export Permit DGLEX-07-2018 MEM Valid from April 25 2018 until April 25 2019
Discharge Abatement Cerro Blanco Project and Environmental Management Plan - Category B2 511-2011/DIGARN/ECM/caml MARN 2028
Property Registry October 31, 2007 2032
Cerro Blanco Building Permit Municipality As. Mita - Difference between previous value and current value must be paid December 29, 2007 Indefinite
Forestry License #1 (East Zone) No. 40-2205-155-1.6-2007 In process of renewal
WTP Handling and disposal of sludge Resolution 00244-2016-DIGARN/FACD/gamc MARN 2028
Amendment Handling and disposal of sludge Resolution 03749-2019 - DIGARN/MOCMD/RJOP 2027
Medical Clinic Sanitary License 14047 Ministry of Health and Social Assistance June 14th, 2016 2026
Resolution: no pre-Hispanic or paleontological remains in the Project area Opinion No. 002/mc.2008 Department of Pre-Hispanic and Colonial Monuments. Indefinite
Diesel Tank Operating License, Own Consumption Lic No. 0627 2029
License for operation and management of Explosives 1942 Undefined
Other resolutions of environmental documents, from previously acquired commitments 2007 Undefined

 

Source: Bluestone, 2019.

 

17.3Water Resources

 

17.3.1Water Quality

 

The environmental monitoring program involves continuous sampling of surface and groundwater. Monitoring at the discharge point shows that all parameters meet the criteria set by MARN and EPA guidelines for ore mining effluent. The study conducted by Consultoria y Tecnologia Ambiental, S.A. (CTA) in 2010 summarizes water quality monitoring results from 2002 to 2007. The findings include:

 

·pH levels between 6.3 and 8.5.

 

·Surface water temperatures between 26°C and 27°C, and groundwater temperatures between 33°C and 72°C.

 

·Conductivity in surface water between 226 and 693 μS/cm, and groundwater between 2,235 and 3,962 μS/cm.

 

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17.3.2Water Management

 

The project's water management infrastructure includes a Water Treatment Plant (WTP), pipelines, settling ponds, channels, ponds, and groundwater wells. Regular monitoring of surface and groundwater has been conducted for the past 10 years. The 2007 Environmental Impact Assessment (EIA) indicates naturally occurring metals like Aluminum, Arsenic, Iron, and Manganese in the region's water.

 

Installed in 2011, the WTP is capable to treat up to 1,500 gallons per minute (gpm) discharging the treated water into Quebrada Tempisque. It removes arsenic using co-precipitation with ferric salt, involving chemical oxidation, pH adjustment, ferric iron addition, and solids separation. Sludge generated from the process is disposed of in lined trenches, as required by the Resolution number 00244-2016- DIGARN/FACD/gamc MARN and will be backfilled into underground facilities at the end of mining operations.

 

A monitoring and sampling program has been in place since 2011, with monthly compliance reports approved by MARN. There have been no incidents of non-compliance, and the program will continue during the project's operational phase.

 

·Surface Water Management: Runoff is classified as “contact” or “non-contact” water. Contact water undergoes treatment before reuse or discharge, while non-contact water is diverted and monitored.

 

·Groundwater Management: Dewatering of the mine is achieved through surface wells and underground sumps. Treated water is either reused or discharged into Quebrada Tempisque. Reinjection wells are used to manage groundwater.

 

17.4Waste Rock and Tailings Management

 

17.4.1Waste Rock

 

The waste rock facility is described in Section 15.20. Temporary storage of waste rock occurs for a maximum of one year before being deposited back underground, as cemented rock fill (CRF) or loose aggregate fill.

 

Based on historical geochemistry test work and waste rock exposure, time will be limited, it was thus assumed that any potential acid generation will not have sufficient time to occur. The current design includes a pond that collects run-off water from the waste rock facility. A water quality monitoring program will be implemented to evaluate groundwater and run-off water quality prior to discharge. The implementation of environmental monitoring and controls will be focused on ensuring that potential acid generation does not occur.

 

17.4.2Tailings

 

Tailings generated from the process plant will be dewatered through filtration before disposal in the Dry Stack Tailings Facility (DSTF), described in Section 15.19.1. A separate contact water management system is designed to collect all run-off from the DSTF to prevent potential surface water contamination. Contact water will be pumped to the WTP for treatment

  

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prior to discharge. The facility is designed to prevent environmental contamination collecting runoff water for treatment.

 

Based on studies of geochemistry characterization and consistent with de approved EIA, tailings are considered non-acid generating (NAG). The next section provides a summary of the geochemical tests performed for the Project.

 

17.4.3Geochemistry Testwork

 

For the Feasibility Study (FS), as carried out in 2019, new tailings sample underwent geochemical testing, and existing data on ore and tailings geochemistry were reviewed.

 

Historical and FS testing results are summarized below:

 

17.4.3.1Previous Test Work

 

A 2006 Water Management Consultants (WMC) report details Preliminary and Phase I characterizations. Waste rock and ore samples from five major ore veins (S1A, S1B, S2, S3, North zone) underwent Acid Base Accounting (ABA) testing (65 samples) and leach extractions (17 samples). Net Neutralization Potential (NNP) values indicated 27 Non-Acid-Generating (NAG) samples, 27 uncertain, and 9 potentially acid generating (PAG). Using Neutralization Potential Ratio -NPR, Phase I data (n=42) showed 48% NAG, 36% PAG and 17 % uncertain samples.

 

Ore samples tested by WMC, which would become tailings, likely retain their original geochemical characteristics post-milling. WMC reported a wide NPR range for ore (0.01 to 1163) with a geometric mean of 3.4, suggesting tailings would likely be NAG.

 

The 2012 DSTF feasibility-level design and cost estimate report by Golder, described geochemical testing for a single tailings sample (run in duplicate) using static and kinetic methods. Results indicated the sample was NAG with sufficient carbonate (calcite) to neutralize any acid from residual sulfide minerals. No evidence of metal leaching was found under aggressive or ambient conditions (SPLP testing). Thus, the evaluated tailings sample showed no potential for acid drainage or metal leaching.

 

17.4.3.2Feasibility Study (2019)

 

In June of 2018, a tailings sample (DS-32-0261 Tailings) was generated by blending various ore types from different locations around the mine and sent to Maxxam Analytics in Burnaby, British Columbia. The sample was subjected to ABA tests, to assess the potential for acid rock drainage, and Shake Flask Extraction (SFE), to assess the potential for metal leaching.

 

The results were similar to historical testing campaigns and showed an abundance of acid neutralizing potential (ANP) compared to the acid generating potential (AGP). With a neutralization potential ratio (NPR) value of 4.8 and a net neutralization potential (NNP) of 41, the tailings were classified as non-acid generating.

 

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Regarding the potential for metal leaching, Mercury is the only constituent leached at a concentration at the limit identified by the International Finance Corporation (IFC). All other metals that might cause impact to the environment if discharged in surface water or groundwater are below both the IFC and limits identified by the U.S. Environmental Protection Agency (EPA).

 

17.5Solid Waste Management

 

17.5.1Non-Hazardous Solid Waste

 

The project proposes continued landfill disposal for inert, non-hazardous solid waste generated during construction and operations. Waste management practices prioritize valorization, recycling, and off-site reuse. The existing landfill will be expanded to accommodate construction waste and remain in use throughout operations, with final reclamation at mine closure. A temporary facility currently stores a Water Treatment Plant (WTP) sludge, which will eventually be disposed of underground. A new temporary storage site will be built near the landfill for ongoing WTP sludge management. The landfill, located on a hilltop outside catchment areas, avoids the need for water diversion. Waste is managed in trenches excavated and compacted by dozers, with each trench backfilled and covered with at least 1.5 meters of material before a new trench is created.

 

17.5.2Solid Hazardous Waste

 

The anticipated hazardous waste primarily includes waste oils, process reagents, and laboratory chemicals. Waste oils will be incinerated or recycled by the supplier. Most reagents and chemicals will be disposed of within the process, with the remainder recycled by the supplier.

 

Cyanide containers and other reagent containers will be washed with fresh water in contained areas, complying with the International Cyanide Code standards. Neutralized products and containers will be disposed of or recycled according to local regulations. Laboratory fire assay wastes, which may contain small amounts of lead, and any lead-contaminated dust will be disposed of in accordance with local regulations. Hazardous material spill clean-ups will be prioritized, involving excavation, neutralization, and disposal of contaminated soils either on-site or at a licensed facility.

 

17.6Flora and Fauna

 

Baseline studies have recorded the region's biodiversity since 2007. Ongoing monitoring indicates minimal impact from the project. The local ecosystem consists of subtropical and tropical dry forests, supporting diverse plant species. Wildlife monitoring shows stable populations of birds, reptiles, and aquatic fauna.

 

Based on a specific Flora and Fauna Management Plan, to protect both flora and fauna environments and species, preventive conservation measures, including habitat relocation for threatened species, such as orchids, tillandsias, cactus or pitayas, have been implemented.

 

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17.7Cultural and Archeological Resources

 

A dedicated onsite team monitors potential impact on cultural and archeological artifacts. Pre-construction inspections and external expert consultations ensure compliance. To date, no significant historical artifacts have been identified within the project’s direct area of influence.

 

17.8Environmental Monitoring

 

The Project is currently operating and reporting on a comprehensive environmental monitoring network consisting of 26 monitoring stations for water quality within and outside of the Project boundaries. There are also nine stations for monitoring air quality.

 

As part of the commitments under the approved EIA, the Project performs monthly monitoring to evaluate the water quality, air quality and noise levels in the direct and indirect areas of influence. Monthly and annual reports are prepared and presented to the Authorities (MARN, Ministerio de Energía y Mina – MEN, Ministerio de Salud de Jutiapa and Ministerio de Ambiente de Jutiapa) to report on the results of monitoring.

 

17.9Environmental Management Plan

 

The Environmental Management Plan (EMP) has been updated using lessons learned from a decade of onsite environmental data collection. The plan is aligned with regulatory requirements and international best practices. It integrates corporate health, safety, and environmental programs, including emergency response strategies.

 

17.10Social Management

 

Aura prioritizes strong community relationships. The Project retains a comprehensive database of community engagement activities and sustainability initiatives.

 

The Social Baseline Study (SBS) included in the EIA was prepared in 2006 using information extracted from 2002 national census. It was updated in 2018 with the most recent Guatemalan official data at that time and information collected from a census conducted in the rural area, along with numerous face-to-face meetings with representatives of local organization.

 

The scope of the study covered the town of Asunción Mita (capital of the municipality), also referred to as the “Urban Area”, plus 6 rural villages included in the external area of influence and nine rural villages located near Lake Güija (collectively referred to as the “Rural Area”). The Lake Güija villages were included in the SBS update to gain a better understanding of the project perception outside of the external area of influence.

 

In Asunción Mita, a total of 28 registered small local community organizations, called Consejos Comunitarios de Desarrollo (COCODEs), manage public budgets and formulate projects to meet community need. During the SBS update, meetings were held with representatives from the 20 active COCODEs in the urban area.

 

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Era Dorada team members are actively involved with local organizations and communities to inform the population project activities and development strategies. In order to respond to the concern around lack of specific knowledge regarding the Project, COCODE representatives and the public were invited to visit the project site, which proved to be very successful.

 

With respect to the surrounding villages, the majority have small populations (i.e. less than 400), except for Trapiche Vargas (566) and San Joaquín (745). The population characteristics of the villages are generally in line with those in the urban area.

 

To manage social issues a dedicated software has been implemented as a tool to register, monitor and report all aspects of the social management program. This has improved access to information for the onsite team and has resulted in more efficient monitoring and reporting activities.

 

The updated Social Management Plan (SMP) incorporates IFC performance standards and includes mechanisms for communication, grievance handling, conjuncture monitoring, local relationship, corporate social alignment, social impact evaluation, continuous training and community involvement. A Social Monitoring Committee (SMC) is in the process of being implemented to ensure transparency.

 

17.11Mine Closure

 

The approved EIA includes a conceptual mine closure plan, which was further refined and is based in a long-term monitoring of water quality and ecosystem restoration, using native plant species in revegetation.

 

Total Closure costs are estimated at US$ 17.9 M.

 

The main requirements of the updated closure plan are summarized below:

 

17.11.1Underground Mine

 

·Progressive underground backfilling of waste rock and tailings.

 

·Removal of all underground equipment.

 

·Portal and vent raises will be blocked with concrete and/or steel plugs.

 

·All infrastructure at portal pads will be removed and concrete pads covered with locally sourced fill and indigenous vegetation.

 

17.11.2Process Plant

 

·Adequate cleaning of infrastructure and drainage of piping before demolition.

 

·Recycling, reuse, and reclamation of materials will be evaluated prior to closure phase to avoid disposal in landfills.

 

·Concrete foundations will remain in place and be covered with locally sourced fill and indigenous vegetation. Surface will be graded to prevent water accumulation.

 

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17.11.3Administration Offices and Ancillary Buildings

 

·Administration offices and all other buildings (including explosives storage warehouse) will be decommissioned and demolished.

 

·Cleaning and decontamination procedures will include equipment/waste management disposal/recycling prior to decommissioning.

 

·Concrete foundations will remain in place and be covered with locally sourced fill and indigenous vegetation. Surface will be graded to prevent water accumulation.

 

17.11.4Dry Stack Tailings Facility (DSTF)

 

·Once mining operations have ceased, the DSTF will be closed, covered with locally sourced fill and revegetated with appropriate native species.

 

·All surface piping, mechanical equipment and electrical services associated with the DSTF will be decommissioned and disposed of.

 

·Soils around the facility will be tested for potential contamination.

 

·Final reclaimed profile will respect the site-specific landform objectives.

 

17.11.5Waste Rock Facility

 

·All waste rock will be hauled and stored underground mine throughout the mining operations.

 

·Natural soil cover will be put in place and vegetation with indigenous species planted over the impacted area.

 

The Dry Stacking Tailings Facility (DSTF) will be constructed continuously over the life of mine using the downstream construction method, so concurrent reclamation will not be possible. At the end of operations, exposed portions of the decant piping will be dismantled and the decant pipes will be plugged below the final surface.

 

The surface of the DSTF will be contoured so that it will shed precipitation, rather than impound it. Topsoil that is stockpiled from the DSTF footprint during construction will be spread over the surface of the DSTF. Native grass seed mixture will be planted to reduce erosion.

 

17.12Potential Risks and Mitigation Actions

 

17.12.1Permitting

 

·Potential Risk: Delays may occur resulting in increased duration of the assumed project development schedule due to new EIAs and permits needed for the power line and approval of permit amendments for injection of mine water and/or new permits.

 

·Mitigation Action: Proceed with application of permit amendments and new EIAs/permits immediately to mitigate potential schedule impact. Continued discussions with local regulatory bodies are required to ensure avoidance of unforeseen delays in permitting.

 

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17.12.2Tailings and Waste Rock

 

·Potential Risk: Tailings and waste rock were assumed to be Non-Acid- Generating (NAG) based on test work completed to date and the limited exposure time at surface for waste rock. Additional test work is required before detail engineering to confirm this assumption. If classification is changed to Potentially-Acid Generating (PAG), the design will need to be updated accordingly, which could result in increased capital costs.

 

·Mitigation Action: The costs of geochemical testing should be included in the project budget and testing should be performed prior to detailed engineering.

 

17.12.3Socio-Political

 

·Socio-Political Risk: Although the local community is favorable to the development of Era Dorada as an underground mine, there is a potential risk of socio-political opposition to mine development which could adversely impact the project development schedule.

 

·Mitigation Action: The development of close relationships with the local communities, landowners and government along with implementation of the Environmental Management Plan (EMP) and Social Management Plan (SMP) is required to engage the people with the project, as well as a consistent monitoring of the main stakeholders.

  

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18CAPITAL AND OPERATING COSTS

 

18.1Capital Cost Estimate

 

LoM Project capital costs total $417 M, consisting of the following distinct phases:

 

·Pre-production capital costs – includes all costs to develop the property to a 1,500 tpd production. Initial capital costs total $264.6 M and are expended over a pre-production period on engineering, construction and commissioning activities.

 

·Sustaining capital costs – includes all costs related to the acquisition, replacement, or major overhaul of assets during the mine life required to sustain operations. Sustaining capital costs total $136.2 M and are expended in operating Years 1 through 16.

 

·Closure Costs – includes all costs related to the closure, reclamation, and ongoing monitoring of the mine, post operations. Closure costs total $17.2 M, and are primarily incurred in Year 1, with costs extending into Year 17 for ongoing monitoring activities.

 

The capital cost estimate was compiled using a combination of database costs, and factors; the overall cost estimate was benchmarked against similar operations. Table 18-1 presents the capital estimate summary for initial and sustaining capital costs with no escalation.

 

18.1.1Capital Cost Summary

 

Table 18-1: Capital cost summary

 

WBS DESCRIPTION Pre-Production Cost
USD ($M)
Sustaining Cost/Closure
USD ($M)
Project Total Cost
USD ($M)
Infrastructure 8.2 8.4 16.6
Power and Electrical 16.7 - 16.7
Water Management 16.1 24.7 40.8
Surface Operations 14.7 1.7 16.4
Mining 63.4 80.2 143.6
Process Plant 49.7 16.7 66.4
Construction Indirect 38.0 4.6 42.6
General Services – Owner’s Costs 21.3 - 21.3
Logistics/ Taxes/ Insurance 9.0 - 9.0
Pre- Production, Start-up & Comissioning 5.0 - 5.0
Contingency 21.9 - 21.9
Closure Costs - 17.2 17.2
TOTAL 263.6 153.5 417.0

 

Source: GE21, 2025.

 

Figure 18-1 and Figure 18-2 present the capital cost distribution for the pre-production and sustaining phases. As typical with underground operations, the majority of sustaining capital costs relate to underground lateral and vertical development. In addition, due to the geothermal nature of the Project, the sustaining capital costs include a significant amount of reinjection well drilling and dewatering wells.

 

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Figure 18-1: Distribution of initial capital cost

 

Source: GE21, 2025.

 

 

Figure 18-2: Distribution of sustaining capital cost

 

Source: GE21, 2025.

 

18.1.2Capital Cost Profile

 

All capital costs for the Project have been distributed against the development schedule in order to support the economic cash flow model. Figure 18-3 presents an annual life of mine capital cost profile including closure years (Year 20-23).

 

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Figure 18-3: Capital cost profile

 

Source: GE21, 2025.

 

18.1.3Key Estimation Assumptions

 

The following key assumptions were made during development of the capital estimate:

 

·Underground development activities will be performed by the Owners forces; and

 

·All surface construction (including earthworks) will be performed by contractors.

 

18.1.4Key Estimation Parameters

 

The following key parameters apply to the capital estimates:

 

·Estimate Class: The capital cost estimates are considered Class 5 estimates (-30%/+50%). The overall Project definition is estimated to be 5%;

 

·Estimate Base Date: The base date of the estimate is May 2025. No escalation has been applied to the capital cost estimate for costs occurring in the future;

 

·Units of Measure: The International System of Units (SI) is used throughout the capital estimate; and

 

·Currency: All capital costs are expressed in US Dollars (US$). Portions of the estimate were estimated in Canadian Dollars (C$) and converted to US Dollars at a rate of CA$1.00: US$0.78

 

18.1.5Basis of Estimate

 

18.1.5.1Mine Capital Cost

 

Capital cost estimates are based on a combination of budgetary quotes from equipment suppliers, in-house cost databases and similar mines in Guatemala. Table 18-2 summarizes the underground mine capital cost estimate.

 

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Table 18-2: Mine capital cost

 

WBS DESCRIPTION Pre-Production Cost
USD ($M)
Sustaining Cost/Closure
USD ($M)
Project Total Cost USD
($M)
Capital Development 20.5 29.7 50.2
Underground Mobile Equipment 0.2 20.6 20.8
Ventilation 3.6 3.0 6.6
Water Management 0.6 0.6 1.3
Fixed Infrastructure 0.6 0.6 1.1
Material Handling 0.6 3.0 3.5
Electrical and Automation 2.6 1.6 4.2
Technical and Safety 1.4 1.2 2.6
Mining - Total 30.1 60.3 90.4

 

Source: GE21, 2025.

 

18.1.5.1.1Capital Development

 

Capital development includes the labour, fuel, equipment usage, power, and consumables costs for lateral and vertical development required for underground access to stopes, and underground infrastructure.

 

· Lateral development fuel, equipment usage, power, and consumables requirements were developed based on the mine plan requirements. Manufacturer database equipment usage rates were applied to the required operating hours; and

 

· Lateral development labour requirements were determined by the required equipment fleet in operation. Supervision and support services were pro-rated to the development costs, based on the mix of underground activities occurring.

 

18.1.5.1.2Underground Mobile Equipment

 

Underground mining equipment quantities and costs were determined through buildup of mine plan quantities and associated equipment utilization requirements. Budgetary quotes were received and applied to the required quantities. The mining fleet is assumed to be provided by a contractor up to the end of Year 2, after which the Owner purchases new mobile equipment and takes control of mining development

 

18.1.5.1.3Underground Infrastructure

 

Design requirements for underground infrastructure were determined from design calculations for ventilation, dewatering, and material handling. Budgetary quotations or database costs were used for major infrastructure components. Allowances have been made for miscellaneous items, such as initial PPE, radios, water supply, refuge stations, and geotechnical investigations. Acquisition of underground infrastructure is timed to support the mine plan requirements.

 

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18.1.5.1.4Capitalized Production Costs

 

Capitalized production costs are defined as mine operating expenses (operating development, mineralized material extraction, mine maintenance, and mine general costs) incurred prior to the introduction of feed to the processing facilities and the commencement of Project revenues. They are included as a pre-production capital cost. Capitalized production costs are included in the asset value of the mine development and are depreciated over the mine life within the financial model.

 

18.1.5.1.5Contractor Development

 

Contractor mine development is based on bid proposals received for Latin American contractors working in similar locations and production rates. Unit cost of development and stoping are based on unit costs for mine activities including drill, blast, muck, haul, and support. A nominal $ 250 k was assumed for contractor mobilization and demobilization, including the cost to set up and take down any additional site facilities to store and maintain equipment. Table 18-3 outlines contractor rates used for capital development.

 

Table 18-3: Contractor capital development rates

 

Activity Profile Unit Price
Drift Rehabilitation 5.0 m x 5.0 m $ 510 per m
Capital Lateral Development 5.0 m x 5.0 m $ 3,080 per m
Capital Vertical Development 4.0 m x 4.0 m $ 1,860 per m
Capital Raisebore 3.0 m diameter $ 2,150 per m

 

Source: GE21, 2025.

 

18.1.5.2Surface Construction Cost

 

Surface construction costs include site development, mineral processing plant, tailings management facility, and on-site and off-site infrastructure. These cost estimates are primarily based on material and equipment costs from material take-offs and detailed equipment lists. Pricing for main equipment and bulk materials was primarily estimated from similar projects, with some factors applied for minor cost elements.

 

Table 18-4 presents a summary basis of estimate for the various commodity types within the surface construction estimates. Growth factors were included above neat material take-off quantities for all areas.

 

Table 18-4: Surface construction basis of estimate

 

Commodity Basis
Access Roads Material take-offs developed based on general arrangements by local contractor.
Bulk Earthworks Model volumes from preliminary 3D grading model. Database unit rates for bulk excavation and fill. Material take-offs for surface drainage, water management ponds and temporary roads from general drawings.
Concrete and Structural Steel Material take-offs estimated and factored costs applied from similar projects.
Buildings and warehouses Buildings according to general arrangements with factored costs for overall building structures.
Mechanical / Electrical Equipment Mechanical/Electrical equipment based on database with factored costs applied from similar projects.

 

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Commodity Basis
Piping Material take-offs are estimated for major pipelines and small-bore piping. Factored costs applied from similar projects.
Dewatering and Injection Wells Number of wells based on dewatering and injection plan. Factored costs applied from similar projects.
Power Transmission Line and Major Sub-stations Estimated based on general arrangements and site layouts. Factored costs applied from similar projects.

 

Source: GE21, 2025.

 

18.1.5.2.1Surface Construction Sustaining Capital

 

Sustaining capital costs include the following dewatering and injection wells.

 

Drystack storage facility earthworks quantities were developed from engineering drawings by the design engineer. Database unit rates benchmarked against projects recently constructed in the region have been applied to the engineered quantities.

 

It is assumed that DSF construction will be performed by a contractor, during both the construction and operations phase.

 

18.1.6Indirect Cost Estimate

 

Indirect costs are classified as costs not directly accountable to a specific cost object. Table 18-5 presents the subjects and basis for the indirect costs within the capital estimate.

 

Table 18-5: Indirect cost basis of estimate

 

Commodity Basis
On Site Contract Services Heavy Lift Crane Services based on estimated durations and historical rates for crane services.
Contractor Field Indirects Estimated by first principles, and including the following items:
- Time-based cost allowance for general construction site services (temporary power, heating and hoarding,  contractor support, etc.) applied against the surface construction schedule
Construction offices and wash car facilities
Safety training, tools and equipment
Environmental cost
Materials Management and Warehouse Operations
Site Maintenance and Temporary Services
Surveying and Quality Assurance
Communications
Contractor facilities and related cost
Construction team facilities, fuel
Freight and Logistics Factor (10%) for freight and logistics related to the materials and equipment required for the crushing plant, mineral processing plant, on-site and off-site infrastructure. Factor excludes mining equipment as prices are FOB site.
Vendor Representatives Estimated by first principles, assessing the equipment supply packages and vendor services hours required for commissioning equipment.
Capital Spares Based on material take-offs from similar projects.
Start-up and Commissioning Included under EPCM (personnel) and Owner's team costs (material and consumables).
First Fills Based on requirements determined by engineering and database pricing.
Detailed Engineering and Procurement Factor applied against direct and indirect hours for engineering management, detailed design, drawings, and major equipment procurement
Project and Construction Management Staffing plan built up against the development schedule for project management, health and safety, construction management, field engineering, project controls, and contract administration. Costs are based on similar project.

 

Source: GE21, 2025.

 

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18.1.7Owner Cost Estimate.

 

Owner’s costs are capitalized in the initial capital costs during the construction phase. Owner’s costs for the project start in Month 10 of Year -2 of the CAPEX cash flows. Any Owner’s costs prior to this are assumed to be within the Owner’s approved budget expenses and are considered sunk costs.

 

18.1.7.1Process Plant Operations

 

The following processing related costs are included in the initial capital:

 

·Management, technical, operations, and maintenance labour employed during the construction phase

 

·First fills of consumables and reagents, and initial consumption during process commissioning to initiate operations, and

 

·Energy costs for power consumed during process commissioning and start-up activities.

 

18.1.7.2Water Treatment Plant Operation

 

The following cost elements are included in the initial capital costs for operation of the water treatment plant:

 

·Technical and operations labour

 

·Maintenance and parts

 

·Power consumption, and

 

·Reagents, consumables, and third-party services.

 

18.1.7.3Dewatering Wells Operations

 

The following costs elements are included in the initial capital costs for operations of dewatering wells:

 

·Supervision, technical and operations labour

 

·Maintenance

 

·Power consumption, and

 

·Reagents, consumables, and third-party services.

 

18.1.7.4Pre-Production G&A – Labour

 

Costs for general and administrative labour are included for the following sectors:

 

·Business Services

 

·General management Sustainability, including:

 

oCommunity relations

 

oHealth and Safety

  

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oEnvironmental

 

·Human resources training

 

·Procurement Logistics

 

·Security, and

 

·Site services and facilities maintenance.

 

Costs associated with the following activities are included within the Sustainability category:

 

·Submit permit amendments for injection of mine water and application for new EIA/permits for access road and powerline, and

 

·Complete land agreement negotiations for access road and power line right-of-ways and injection well pads.

 

18.1.7.5Pre-Production G&A – Equipment

 

Costs for owner site support equipment usage are included for the following sectors:

 

·Site Services

 

·Warehouse / material management

 

·Security

 

·Health, Safety, and Environment, and

 

·Administration / management.

 

18.1.7.6Pre-Production G&A – Expenses and Service

 

Costs for general and administrative expenses and fees are included for the following sectors:

 

·Health, safety and medical supplies

 

·Staff safety equipment

 

·Environmental services, fees, and outside laboratory costs

 

·Human resources (training, recruitment)

 

·Construction insurance

 

·Community relations and programs

 

·Legal and regulatory, including property tax

 

·External consulting IT and communications

 

·Site office costs

 

·Office lease and services for Guatemala City

 

·Waste disposal, and

 

·Existing infrastructure power and maintenance.

 

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18.1.8Closure Cost Estimate

 

Closure costs have been estimated based on the typical closure, reclamation, and monitoring activities for an underground mine. Activities include:

 

·Removal of all surface infrastructure and buildings

 

·Closure and capping of the DSTF

 

·Removal of all fixed underground equipment

 

·Closure of the underground mine portals

 

·Power transmission line and substation removal

 

·Re-vegetation and seeding, and

 

·Ongoing site monitoring.

 

The majority of closure costs are incurred immediately following completion of operations (20). Monitoring activities are anticipated to extend to Year 24. Table 18-6 provides a summary of closure cost categories.

 

Table 18-6: Closure estimate summary

 

Item Estimated Cost (M$)
Safety 0.64
Underground Mine 1.36
Infrastructure 0.46
Process Plant 6.78
Water Treatment Plant 1.25
Piping, Ponds and Tanks 2.12
Switchyard and Power Distribution 0.61
Administration Offices and Ancillary Buildings 0.51
Drystack Tailings Facility (DSTF) 2.74
Waste Rock Dump 0.43
Wells 1.84
Indirect Costs 6.14
Monitoring 2.20
Total Closure 27.09

 

Source: GE21, 2025.

 

18.1.9Contingency

 

Contingency has been applied to the estimate on a line-by-line basis as a deterministic allowance by assessing the level of confidence of the scope definition, supply cost and installation cost, and then applying an appropriate weighting to each of the three inputs. The overall recommended pre-production contingency resulted in approximately 12% of direct, indirect, and Owner’s costs.

 

18.1.10Capital Estimate Exclusions

 

·The following items have been excluded from the capital cost estimate:

 

·Working capital (included in the financial model)

 

·Financing costs

 

·Currency fluctuations

 

·Lost time due to severe weather conditions beyond those expected in the region

 

·Lost time due to force majeure

  

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·Additional costs for accelerated or decelerated deliveries of equipment, materials or services resultant from a change in Project schedule

 

·Warehouse inventories, other than those supplied in initial fills, capital spares, or commissioning spares

 

·Any Project sunk costs (studies, exploration programs, etc.)

 

·Value added tax (VAT)

 

·Closure bonding, and

 

·Escalation cost.

 

18.2Operating Cost Estimate

 

The operating cost estimate in this study includes the costs to mine and process the mineralized material to produce doré, along with site services to maintain the site, and general and administrative expenses (G&A). These items total the Project operating costs and are summarized in Table 18-1. The target accuracy of the operating cost is -30/+50%. The operating cost estimate is broken into four major sections:

 

·Underground mining

 

·Processing

 

·Site Services, and

 

·General and Administrative (G&A).

 

The total operating unit cost is estimated to be US$ 117.78/t processed. Average annual, total LOM and unit operating cost estimates are summarized in Table 18-7. The unit rates in this table include tonnes mined during pre-production. Figure 18-4 illustrates the operating cost distribution.

 

Operating costs are expressed in US dollars. No allowance for inflation has been applied.

 

Table 18-7: Breakdown of Estimated Operating Costs

 

Operating Costs Avg Annual (M$) $/t processed LOM (M$)
Mining 38.70 100 890.01
Processing 12.38 32 284.80
Site Services 6.97 18 160.20
G&A 7.74 20 178.00
Total 65.78 170 1 513.01

 

Source: GE21, 2025.

 

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Figure 18-4: Operating Cost Distribution

 

Source: GE21, 2025.

 

The main operating cost component assumptions are shown in Table 18-8.

 

Table 18-8: Main OPEX Component Assumptions

 

Item Unit Value
Electrical Power Cost $/kWh 0.06
Overall Power Consumption (all facilities) kWh/t processed 215
Diesel Cost (delivered) $/litre 0.79
LOM Average Operations Workforce employees 584

 

Source: Bluestone, 2019.

 

18.2.1Operation Labour

 

This section provides an overview of total workforce and the methods used to compile the labour rates. Table 18-9 summarizes the total planned workforce during Project operations.

 

Table 18-9: Main OPEX Component Assumptions

 

Operating Costs Construction Operations
Mining 193 329
Processing 97 97
Site Services 64 69
G&A 120 105
Total 474 600

 

Source: Bluestone, 2019.

 

Labour is a significant portion of annual operating cost. Labour rates include base wage and allowances for overtime, night shift, insurance, tax, and benefits. Labour burdens were assembled using first principles and range from 34 to 42%. The following items are included in the burdened labour rates:

 

·Unscheduled Overtime at 10%;

 

·Social security at 12.67%;

 

·Yearly Christmas bonus at 1 month salary per year;

 

·Yearly Bonus 14 at 1 month salary per year;

 

·Vacation pay at 4%;

 

·Statutory Holiday pay at 3%;

 

·Yearly Insurance Payments of $ 54.00/employee/month; and

 

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·Monthly savings funds at 13%.

 

18.2.2Mining Operating Cost Estimate

 

Mining operating costs listed in Table 18-10 are averaged over the life of mine. Mine operating costs have been built up from using a combination of first principle engineering and equivalent Project scaling.

 

Mine operating unit costs are summarized below in Table 18-10 and Figure 18.3 and include:

 

·Waste development – costs related to the drilling, blasting, mucking, and hauling of non-capital development

 

·Production – costs related to the drilling, blasting, mucking, and hauling of ore for both longhole and cut and fill stoping

 

·Backfill – costs related to CRF and paste backfill operations, including the CRF and paste plants

 

·Mine Maintenance – maintenance labour costs that support all other sectors, and

 

·Mine General – costs related to mine support activities, such as technical services, shared infrastructure, support equipment, and definition drilling.

 

Table 18-10: Underground Mine Operating Costs

 

Mining Category Unit Cost ($/t processed) LOM Cost (M$)
Lateral Waste Development 6.27 55.78
Lonhole Stoping 21.95 195.37
Cut and Fill Stoping 25.52 227.12
Backfill 24.92 221.80
Mine Maintenance 3.88 34.53
Mine General 17.46 155.40
Total 100.00 890.01

 

Source: GE21, 2025.

 

18.2.3Processing Operating Cost

 

Process operating costs were developed using labour rates based on operating mines in the area and sufficient personnel to operate the process plant, factored maintenance cost, budget quotes for consumables and a factored power requirement. Process operating costs are summarized below in Table 18-11. Costs are subdivided into operating categories.

 

Table 18-11: Process Operating Costs

 

Mining Category Unit Cost ($/t processed) LOM Cost (M$)
Labour 10.53 93.69
Power 6.81 60.59
Maintenance and Consumables 14.67 130.53
Total Processing OPEX 32.00 284.80

 

Source: GE21, 2025.

 

Process labour includes burden for salaried and hourly employees to account for in-country benefits, training, production bonus and potential ex-patriot benefits & costs.

 

Equipment maintenance was calculated by applying a factor of 4% to major process equipment cost.

 

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Power costs were calculated from the total installed power assuming $ 0.06/kWh.

 

18.2.4General and Administration Operating Cost Estimate

 

General and Administration costs include all on-site activities including but not limited to dry stack tailings (DST) haulage, personal protection services, water treatment plant operation, site services equipment and labour, office operating costs and associated labour. The summary of costs is shown in Table 18-12, averaged over the life of mine.

 

Table 18-12: General and administration (G&A) operating cost summary

 

G&A Labour Total  4.18  100.20  11.26
Nusiness Services  0.85  20.43  2.30
General Management  0.50  12.11  1.36
Sustainability  1.26  30.27  3.40
Human Resources  0.27  6.51  0.73
Purchasing and Logistics  0.20  4.84  0.54
Security  0.52  12.56  1.41
G&A Service Staff  0.30  7.27  0.82
Project Team  0.26  6.21  0.70
G&A Services and Expenses Total  3.24  77.80  8.74
Accommodations  0.18  4.24  0.48
Heath and Safety, Medical, and First Aid  0.22  5.30  0.60
Environmental  0.44  10.60  1.19
Human Resources  0.63  15.14  1.70
Insurance and Legal  0.84  20.13  2.26
External Consulting  0.47  11.20  1.26
IT and Communications  0.16  3.78  0.43
Office and Miscellaneous Costs  0.01  0.30  0.03
Satellite Office  0.11  2.57  0.29
Employee Travel  0.20  4.84  0.54
Total G&A  7.42  178.00  20.00

 

Source: GE21, 2025.

 

18.2.4.1G&A Labour Requirements

 

Table 18-13 lists the site supervision and support personnel requirements and costs.

 

Table 18-13: G&A Labour Requirements & Costs

 

Labour Salary/Hourly Loaded Pay (US$) Quantity
Mine/General Manager Salary 243,459 1
Site Administrator Salary 12,956 1
Accounting & Taxes Manager Salary 20.34 1
Accounting/Payroll Coordinator Salary 15,417 2
Human Resources Manager Salary 22.802 1
Human Resources Clerk Salary 12,956 1
Trainer Salary 15,417 2
Community Relations Manager Salary 74,493 1
Community Relations Coordinator Salary 15,417 1
IT/Telecom. Technician Salary 22.802 1
Procurement/Contracts/Logistics Manager Salary 25.263 1
Procurement/Contracts Agent Salary 20.34 1
Warehouse Operators Salary 15,417 1
Multi-Equipment Operators Salary 15,417 2
Health, Safety, and Security Manager Salary 49.878 1
Health & Safety Coordinator Salary 22,802 2
First Aid Attendant/Nurse Salary 7,591 4
Environmental Manager Salary 49,878 1
Environmental Technician Salary 15.417 1
Environmental Coordinator Salary 25,263 2
Protective Services Officials Salary 12,956 30
Site Services Foreman Salary 37.571 2
Carpenters Salary 15.417 1
WTP Operator Salary 15.417 4

 

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Labour Salary/Hourly Loaded Pay (US$) Quantity
Multi-Equipment Operatior Salary 15,417 4
Skilled Labourers Salary 11,725 2
Total G&A Labour     71

 

Source: Bluestone, 2019.

 

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19ECONOMIC ANALYSIS

 

The economic analysis for the Project is based on Mineral Resource estimates, including the annual mine production schedule previously outlined in this report. As required under SK-1300, the results of this analysis should not be interpreted as demonstrating the economic viability of the project.

 

The outcome of the economic analysis is subject to known and unknown risks, uncertainties, and other factors that may cause actual results to differ materially from them. The information on which this analysis is based is listed below:

 

·Mineral Resource Estimates

 

·Assumed fixed exchange rate

 

·Proposed mine production plan

 

·Projected mining and processing recovery rates

 

·Fixed installed processing plant capacity

 

·Assumptions on closure costs

 

·Assumptions on environmental, licensing, and social risks

 

·Changes in production costs relative to the assumptions

 

This analysis does not rely on:

 

·Unrecognized environmental risks

 

·Unanticipated recovery expenses

 

·Different geotechnical and/or hydrogeological considerations during mining

 

·Unexpected variations in the quantity of mineralized material, grade, metallurgical recovery efficiency, and plant recovery efficiency

 

·Accidents, labour disputes, and other mining industry risks

 

·Changes in tax rates

 

·Assumptions of commercial discounts that are not foreseen in the financial analysis

 

Based on the Cash Flow Model results, the Era Dorada Project has a project-level after-tax NPV of US$ 485.5 million at a 5% discount rate, an after-tax IRR of 23.8%, and a payback period of 3.75 years. Project results are summarized in Table 19-1.

 

Table 19-1: Summary of key financial results

 

Description Unit Value
Price Au1 $/oz  $        2.409,68
Price Ag1 $/oz  $             28,64
Au recovery kozt  $        1.377,14
Ag recovery kozt  $        4.308,07
Gold Equivalent - GEO (kozt) kozt  $        1.426,88
NPV  After-Tax $MUS$               485,49
IRR - After tax % 23,8%
After-tax Payback2 years                   3,75
NPV  Pre-Tax @5% $MUS$               680,39
IRR - Pre tax % 30,4%
Pre-Tax Payback2 years                   3,35
Life of Mine (LOM) years  $             17,00

 

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Description Unit Value
Average Annual Production (Au) kozt  $             81,01
Average Annual Production (Ag) kozt  $           253,42
Average Annual Production (Au eq) kozt  $             83,93
Total Oz Payable Au kozt  $        1.376,04
Total Oz Payable Ag kozt  $        4.286,53
Selling Cost (MUS$) $MUS$  $             17,19
Royalties $MUS$  $           242,07
Opex $MUS$  $        1.512,99
Initial Capital $MUS$  $           263,55
Sustaining and Closure Capital $MUS$  $           153,41
Average Cash Cost $MUS$  $        1.072,40
Average All in Sustaining Cost $/oz  $        1.179,91

 

Notes:

1.Price resulting from the weighted average of forecast gold and silver prices from 2028 onwards

2.Payback period post-construction

Source: GE21, 2025.

 

19.1Methodology

 

An economic model was developed to estimate the Project’s post-tax annual cash flow and sensitivity analysis based on an assumed discount rate of 5%. Capital and operating cost estimates were summarized in Section 18 of this report. The economic analysis was performed without inflation.

 

This analysis was conducted with the following assumptions:

 

·Year -1 corresponds to pre-production phase

 

·Price inflation and escalation factors are ignored (constant-dollar basis)

 

·Results are based on 100% equity capital

 

·Project revenue is derived from selling a basket list of graphite production

 

·All production is sold at the year of production

 

19.2Gold and Silver Prices

 

The silver and gold prices used for the economic evaluation are 28,44 US$/oz and 2,389 US$/oz respectively based on the long term consensus forecast from over 20 investment banks.

 

19.3Mine Production

 

The annual production figures were derived from the mine plan presented in Section 13. Over the life of the mine, a total of 9 Mt of ore is mined.

 

19.4Plant Production

 

The silver and gold production plan was estimated considering the recoveries described in Sections 14. The process plant is designed based on the following criteria:

 

·1,500 tpd ore production;

 

·85% average silver recovery;

 

·96 % average gold recovery;

 

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·Estimated life of mine gold production of 1.4 million ounces and silver production of 4.3 million ounces.

 

19.5Revenue

 

Annual revenue is calculated by applying the estimated silver and gold prices to the annual payable metal for each operating year. Gross revenue represents the total value of payable silver and gold. Net Smelter Return (NSR) revenue accounts for associated selling costs, while Net Revenue further accounts for royalties payable, as shown on Table 19-2.

 

Table 19-2: Revenue composition

 

Description LOM (MUS$)
Gold Gross Revenue  $ 3,315.80
Silver Gross Revenue  $ 122.78
Total Gross Revenues  $ 3,438.59
Smelting and Refining  $ 17.19
Total NSR Revenues  $ 3,421.39
Goldcorp Royalty  $ 35.92
Newmont Royalty  $ 34.21
Guatemalan Government Royalty  $ 171.93
Total Net Revenues  $ 3,179.33

 

Source: GE21, 2025.

 

19.6Total Operating Cost

 

The average total unit operating cost over the life of mine is estimated at $170 t of ore processed. A detail of the operating cost is shown in Section 18 and summarized in Table 19-3.

 

Table 19-3: Detailed operating costs

 

Description LOM (MUS$) $/t ROM $/ozt GEO
Mining  $ 890.00  $ 100.00  $ 623.73
Processing  $ 284.80  $ 32.00  $ 199.59
Site Service  $ 160.20  $ 18.00  $ 112.27
SG&A  $ 178.00  $ 20.00  $ 124.75
Total Operating Costs  $ 1,512.99  $ 170.00  $ 1,060.35

 

Source: GE21, 2025.

 

19.7Royalty Rights

 

Royalties in mining are financial compensation paid to the State for the right to exploit mineral resources, contributing to the redistribution of the benefits of mining activities. In Guatemala, the royalties due for mineral extraction amount to 5% of gross revenue. Additionally, a 1.05% rate on the Net Smelter Return is considered for royalty payments to Goldcorp Royalty.

 

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19.8Capital Expenditure

 

19.8.1Initial Capital

 

The financial analysis for the Era Dorada Project was prepared under the assumption of 100% equity financing for initial capital expenditures. The total initial capital cost estimate is USD 263.5 million and includes expenditures related to:

 

·pre-stripping;

 

·power supply and electrical infrastructure;

 

·water management systems;

 

·surface infrastructure;

 

·mine equipment and development;

 

·process plant construction;

 

·indirect costs;

 

·general services;

 

·owner's costs, logistics, taxes, and insurance.

 

The estimate also includes costs associated with process plant operations during pre-production, start-up, and commissioning, as well as an appropriate contingency allowance.

 

19.8.2Sustaining capital

 

Sustaining capital is estimated at 136 MUSD and includes the renewal of the mining fleet, wtaer management, process plant maintenance, surface operations and indirect constructions and infraestructure costs.

 

19.8.3Remediation and Closure Capital

 

The project includes a provision of USD 17.19 million, to be accumulated over the life of mine through the allocation of 0.5% of gross revenue. This amount is considered appropriate given the scale of the project.

 

19.9Total All in Sustaining Cost

 

The average total All-In Sustaining cost over the life of the mine is estimated at $1,179 per ounce of payable gold equivalent, as shown in Table 19-4.

 

Table 19-4: All in sustaining costs composition

 

Description LOM (MUS$) $/t ROM $/ozt GEO
Mining  $                   890.00  $      100.00  $            623.73
Processing  $                   284.80  $         32.00  $            199.59
Site Service  $                   160.20  $         18.00  $            112.27
SG&A  $                   178.00  $         20.00  $            124.75
Selling Costs and Royalies  $                      17.19  $           1.93  $               12.05
Sustaining and closure capital  $                   153.41  $         17.24  $            107.51

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Description LOM (MUS$) $/t ROM $/ozt GEO
Total All in Sustaining Costs  $                1,683.59  $      189.17  $         1,179.91

 

Source: GE21, 2025

 

19.10Working Capital

 

For the purposes of estimating working capital requirements, the following assumptions were adopted for the financial analysis in Table 19-5.

 

Table 19-5: Working capital periods

 

Working Capital Component Days
Average collection period 30
Average inventory turnover period 30
Average payment period 30

 

Source: GE21, 2025.

 

19.11Depreciation

 

Depreciation of capital assets has been estimated at 10% annually for the purpose of simplifying the analysis. Fiscal reserves related to exploration and resource development have been excluded from the depreciation calculation. No salvage value has been applied to capital items, as any salvage proceeds are treated as taxable income.

 

19.12Exchange Rate Forecast

 

The exchange rate was defined based on parameters adopted in international projects, not using values projected by any financial institution. The exchange rate used was Q7.75/US$.

 

19.12.1Income Tax

 

Income tax applies to the profits earned by companies and other legal entities. It is calculated based on the accounting results determined at the end of a reporting period, such as a quarter or a fiscal year. In Guatemala, companies are taxed on their gross profits at a rate of 25% of taxable income.

 

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19.13

Discounted Cash Flow

  

A simplified discounted cash flow was developed to evaluate the Project based on economic-financial parameters, mining schedule results, and capital, sustaining and operating cost estimates, and is presented in Table 19-6.

 

Table 19-6 - Simplified Discounted Cash Flow

 

Cash Flow Un.  Total  -2 -1 1 2 3 4 5 6 7 8 9 10  
Gold Equivalent Ounce ktoz 1,426.88 - - 97.91 90.25 88.25 86.39 80.50 95.15 81.77 87.20 84.48 81.46  
Price  Au US$/toz 2,409.68 - - 2,693.00 2,389.00 2,389.00 2,389.00 2,389.00 2,389.00 2,389.00 2,389.00 2,389.00 2,389.00  
Gross Revenue M US$ 3,438.59 - - 263.67 215.60 210.82 206.38 192.30 227.30 195.35 208.33 201.81 194.61  
Freight & Refining M US$ 17.19 - - 1.32 1.08 1.05 1.03 0.96 1.14 0.98 1.04 1.01 0.97  
Net Smelter Return M US$ 3,421.39 - - 262.35 214.52 209.77 205.35 191.34 226.17 194.37 207.29 200.81 193.64  
Royalties M US$ 242.07 - - 18.56 15.18 14.84 14.53 13.54 16.00 13.75 14.67 14.21 13.70  
Tax Sales M US$ - - - - - - - - - - - - -  
Net Revenue M US$ 3,179.33 - - 243.79 199.35 194.92 190.82 177.81 210.16 180.62 192.62 186.60 179.94  
Total opex M US$ 1,512.99 - - 93.08 93.08 93.08 93.08 93.08 93.08 93.08 93.08 93.08 93.08  
Cash Cost US$/oz 1,072.40 - - 964.08 1,043.27 1,066.67 1,089.36 1,168.22 990.18 1,150.21 1,079.29 1,113.73 1,154.49  
All in Sustaining Costs US$/oz 1,179.91 - - 1,215.77 1,351.83 1,239.35 1,165.06 1,291.94 1,041.61 1,212.88 1,145.06 1,190.31 1,242.17  
Gross Profit M US$ 1,666.33 - - 150.72 106.27 101.85 97.74 84.73 117.09 87.54 99.54 93.52 86.87  
Depreciation M US$ 373.38 - - 26.35 28.82 31.60 33.13 33.78 34.78 35.27 35.78 36.35 26.46  
EBIT (US$) M US$ 1,292.95 - - 124.36 77.45 70.25 64.62 50.95 82.31 52.28 63.77 57.17 60.41  
Income Tax M US$ 323.24 - - 31.09 19.36 17.56 16.15 12.74 20.58 13.07 15.94 14.29 15.10  
Operational profit(US$) M US$ 969.72 - - 93.27 58.09 52.68 48.46 38.21 61.73 39.21 47.82 42.88 45.31  
(=) EBIT M US$ 1,292.95 - - 124.36 77.45 70.25 64.62 50.95 82.31 52.28 63.77 57.17 60.41  
Depreciation M US$ 373.38 - - 26.35 28.82 31.60 33.13 33.78 34.78 35.27 35.78 36.35 26.46  
(=) EBITDA M US$ 1,666.33 - - 150.72 106.27 101.85 97.74 84.73 117.09 87.54 99.54 93.52 86.87  
(-) Capex M US$ 399.76 105.42 158.13 24.64 27.85 15.24 6.54 9.96 4.89 5.12 5.74 6.47 7.14  
(+) Residual Value M US$ 26.38 - - - - - - - - - - - -  
(+-) Working Capital M US$ - - - 22.01 0.38 -0.22 -3.33 -0.80 2.04 -1.84 0.76 -0.36 2.48  
(-)ARO M US$ 17.19 - - - - - - - - - - - -  
(-) Income Tax M US$ 323.24 - - 31.09 19.36 17.56 16.15 12.74 20.58 13.07 15.94 14.29 15.10  
(=) Post-tax Cash Flow M US$ 952.52 -105.42 -158.13 72.97 58.69 69.27 78.38 62.83 89.58 71.19 77.11 73.13 62.14  
(=) Post-Tax Acumulated Cash Flow M US$ 7,345.55 -105.42 -263.55 -190.58 -131.89 -62.63 15.75 78.59 168.16 239.35 316.45 389.58 451.72  
(=) Pre-Tax Cash Flow M US$ 1,275.76 -105.42 -158.13 104.06 78.05 86.83 94.53 75.57 110.15 84.26 93.05 87.42 77.24  
(=) Pre-tax Acumulated Cash Flow M US$ 10,578.08 -105.42 -263.55 -159.49 -81.44 5.39 99.92 175.49 285.64 369.90 462.95 550.37 627.61  

 

Source: GE21, 2025.

 

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Table 19-7 - Simplified Discounted Cash Flow (continued)

 

Cash Flow Un.  Total  11 12 13 14 15 16 17 18  
Gold Equivalent  Ounce ktoz 1,426.88 80.52 81.58 90.87 94.61 94.60 89.48 21.88    
Price  Au US$/toz 2,409.68 2,389.00 2,389.00 2,389.00 2,389.00 2,389.00 2,389.00 2,389.00    
Gross Revenue M US$ 3,438.59 192.36 194.88 217.08 226.03 225.99 213.78 52.28    
Freight & Refining M US$ 17.19 0.96 0.97 1.09 1.13 1.13 1.07 0.26    
Net Smelter Return M US$ 3,421.39 191.40 193.91 216.00 224.90 224.86 212.71 52.01    
Royalties M US$ 242.07 13.54 13.72 15.28 15.91 15.91 15.05 3.68    
Tax Sales M US$ - - - - - - - -    
Net Revenue M US$ 3,179.33 177.86 180.19 200.72 208.98 208.95 197.66 48.33    
Total opex M US$ 1,512.99 93.08 93.08 93.08 93.08 93.08 93.08 23.79    
Cash Cost US$/oz 1,072.40 1,167.86 1,152.91 1,036.23 995.71 995.86 1,052.07 1,099.24    
All in Sustaining Costs US$/oz 1,179.91 1,257.62 1,242.06 1,085.75 1,028.60 1,061.24 1,117.37 1,358.52    
Gross Profit M US$ 1,666.33 84.79 87.12 107.64 115.91 115.88 104.59 24.54    
Depreciation M US$ 373.38 11.36 9.62 7.56 6.49 6.14 5.18 4.71    
EBIT (US$) M US$ 1,292.95 73.43 77.50 100.08 109.42 109.73 99.40 19.83    
Income Tax M US$ 323.24 18.36 19.37 25.02 27.36 27.43 24.85 4.96    
Operational profit(US$) M US$ 969.72 55.07 58.12 75.06 82.07 82.30 74.55 14.88    
(=) EBIT M US$ 1,292.95 73.43 77.50 100.08 109.42 109.73 99.40 19.83    
Depreciation M US$ 373.38 11.36 9.62 7.56 6.49 6.14 5.18 4.71    
(=) EBITDA M US$ 1,666.33 84.79 87.12 107.64 115.91 115.88 104.59 24.54    
(-) Capex M US$ 399.76 7.23 7.27 4.50 3.11 0.34 0.17 -    
(+) Residual Value M US$ 26.38 - - - - - - 26.38    
(+-) Working Capital M US$ - -0.44 0.11 1.24 0.49 -0.01 -0.73 -16.29 -5.50  
(-)ARO M US$ 17.19 - - - - 5.85 5.67 5.67 -  
(-) Income Tax M US$ 323.24 18.36 19.37 25.02 27.36 27.43 24.85 4.96 -  
(=) Post-tax Cash Flow M US$ 952.52 59.65 60.36 76.88 84.95 82.27 74.62 56.58 5.50  
(=) Post-Tax Acumulated Cash Flow M US$ 7,345.55 511.37 571.73 648.61 733.55 815.82 890.44 947.02 952.52  
(=) Pre-Tax Cash Flow M US$ 1,275.76 78.00 79.73 101.90 112.30 109.70 99.47 61.54 5.50  
(=) Pre-tax Acumulated Cash Flow M US$ 10,578.08 705.61 785.35 887.25 999.55 1,109.25 1,208.72 1,270.26 1,275.76  

 

Source: GE21, 2025.

 

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The economic results are presented in Table 19-8.

 

Table 19-8: Simplified Discounted Cash Flow Results

 

Discount Rate (%) 5%
NPV – After Tax (M US$) 485.49
IRR – After Tax (%) 23.8%
Payback - After Tax (years) 3.75

 

Source: GE21, 2025.

 

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20ADJACENT PROPERTIES

 

There are no adjacent properties relevant to the scope of this report.

 

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21OTHER RELEVANT DATA AND INFORMATION

 

There is no other relevant data or information relative to the scope of this report.

 

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22INTERPRETATION AND CONCLUSIONS

 

22.1Geology & Mineral Resources

 

Era Dorada is a classic hot springs-related, low-sulfidation epithermal gold-silver deposit comprising both high-grade vein and low-grade disseminated mineralization. Most of the high-grade mineralization is hosted in the Mita unit as two upward-flaring vein swarms (north and south zones) that converge downwards and merge into basal feeder veins where drilling has demonstrated widths of high-grade mineralization (e.g., 15.5 m 21.4 g Au/t and 52 g Ag/t). Bonanza gold grades are associated with ginguru banding and carbonate replacement textures. Sulfide contents are low, typically <3% by volume. Low-grade disseminated and veinlet mineralization in wall rocks around the high-grade veins is well documented in drilling since discovery of the deposit, with grades typically ranging from 0.3 to 3.0 g/t Au.

 

The Mita rocks are overlain by the Salinas unit, a sub-horizontal sequence of volcanogenic sediments and sinter horizons approximately 100 m thick that form the low-lying hill at the Project. Low-grade disseminated and veinlet mineralization within and as halos around the high-grade vein swarms is well documented in drilling since discovery of the deposit, with grades typically ranging from 0.3 to 1.5 g Au/t. The overlying Salinas cap rocks are also host to low-grade mineralization associated with silicified conglomerates and rhyolite intrusion breccias.

 

Mineral exploration activities performed at Era Dorada have been performed in accordance with “CIM Mineral Exploration Best Practice Guidelines” dated November 23, 2018.

 

The mineral resource has a footprint of 800 x 400 m between elevations of 525 and 200 masl. The mineral resource estimate is the result of 153,003 m of drilling by Bluestone and previous operators (totalling 1,256 drill holes and channel samples). There are 130,307 gold assays which average 0.68 g/t and 130,238 silver assays or 153,003 m total which average 3.75 g/t. Bulk densities were assigned to individual rock types and assigned on a block-by-block basis using measurement data by lithology and mineralized vein.

 

The 3.4 km of underground infrastructure allowed for underground mapping, sampling, and over 30,000 m of underground drilling enhanced the understanding and validation of the Era Dorada geological model. The mineral resource estimate included an estimate of dilutive material, some of which has proven to be economic and have a reasonable prospect of economic extraction. Therefore, improved and refined geological models of the lithological units was required. These broad mineralized lithologies are host to the high-grade veins that have been the focus of the potential underground mining scenario. The resulting domain models and estimation strategy were designed to accurately represent the grade distribution.

 

The estimate was completed using MineSightTM software using a 3D block model (5 m by 5 m by 5 m). Interpolation parameters have been derived based on geostatistical analyses conducted on 1.5-meter composited drill holes. Block grades have been estimated using ordinary kriging (OK) methodology and the mineral resources have been classified based on proximity to

 

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sample data and the continuity of mineralization in accordance with CIM’s “Definition Standards for Mineral Resources and Mineral Reserves” dated May 19, 2014, and “CIM Estimation of Mineral Resources and Mineral Reserves Best Practice Guidelines” dated November 29, 2019.The mineral resources are presented in at a 2.25 g/t Au/t cut-off grade.

 

Table 22-1: Resource Estimate using 2.25 g Au/t Cut-off

 

Resource Category Tonnes (kt) Au Grade (g/t) Ag Grade (g/t) Contained Gold (koz) Contained Silver (koz)
Measured          
Indicated 6,349 9.31 31.54 1,901 6,439
Measured & Indicated 6,349 9.31 31.54 1,901 6,439
Inferred 605 6.02 19.68 117 383

 

Notes:

The mineral resource statement is subject to the following:

1.Mineral Resources are reported in in accordance with S-K 1300.

2.Mineral resource estimates have been prepared by Garth Kirkham, P.Geo., a Qualified Person as defined by SK-1300.

3.The Mineral Resource estimate is reported on a 100% ownership basis.

4.Underground mineral resources are reported at a cut-off grade of 2.25 g Au/t. Cut-off grades are based on a assumed metal prices of US$ 2,500/oz gold and US$ 28/oz silver, and assumed metallurgical recovery, mining, processing, and G&A costs.

5.Mineral Resources are reported without applying mining dilution, mining losses, or process losses.

6.Resources are constrained within underground shapes based on reasonable prospects of economic extraction, in accordance with SK-1300. Reasonable prospects for economic extraction were met by applying mining shapes with a minimum mining width of 2.0 m, ensuring grade continuity above the cut-off value, and by excluding non-mineable material prior to reporting.

7.Metallurgical recoveries reported as the average over the life of mine and are assumed to be 96% Au and 85% Ag, respectively.

8.Bulk density is estimated by lithology and averages 2.47, 2.57 and 2.54 g/cm3 for the Salinas, Mita and mineralized vein domains, respectively.

9.Mineral resources are classified as Indicated, and Inferred based on geological confidence and continuity, spacing of drill holes, and data quality.

10.Effective date of the mineral resource estimate is December 31, 2024.

11.Tonnage, grade, and contained metal values have been rounded. Totals may not sum due to rounding.

12.Mineral resources are not mineral reserves and do not have demonstrated economic viability.

Source: Kirkham, 2025.

 

In addition, there has been mixed grade material mined during the creation of the extensive, existing ramp network which has been stockpiled adjacent to North Ramp entrance. Table 22-2 shows the volume and tonnage based on an unconsolidated specific gravity of 2.0 gm/cm3 along with gold and silver grades and metal content. These resources are classified as measured.

 

Table 22-2: Stockpile Resource Estimate (Measured Resource)

 

Volume (BCM) Mine (t) Au (g/t) Ag (g/t) Au (oz) Ag (oz)
14,863 29,726 5.35 22.59 5,108 21,590

 

Source: Kirkham, 2019.

 

22.1.1Risks

 

The most significant project risks are summarized below:

 

·Commodity Prices (Gold, Silver) – Lower commodity prices will change the size and grade of the potential targets. Conversely, increased commodity prices will improve economics and resources.

 

·Although there is a relatively high degree of confidence related to geological continuity and grade variability, vein models and grade distributions may adjust with further data and structural interpretations.

 

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22.2Mineral Processing and Metallurgical Testing and Processing and Recovery Methods

 

The metallurgical test work campaigns resulted in an adequate database for estimating overall gold and silver recoveries. The same campaigns adequately supported the selected flowsheet configuration, the latter strictly following standard practices of similar industrial circuits.

 

The stipulated capacity of the processing plant for the Project is 1000 tpd, for a ground product with a P80 of 0.053 mm.

 

The selected method for cyanide neutralization (SO2/air) resulted in adequate performance according to environmental regulatory specifications.

 

22.3Mining Methods, Infrastructure, Capital and Operating Costs

 

The Project outlines a conceptual mine plan involving the extraction of 8.9 Mt of ROM over a 17-year LoM, with a production rate of 1,500 tpd. The selected underground mining methods is suitable for ensuring stable and consistent mill feed throughout the mine life.

 

The project features a comprehensive and integrated infrastructure plan that includes new access roads, power supply systems, water management facilities, a process plant, and tailings and waste rock storage. Existing support infrastructure will be leveraged, while new installations will address essential gaps in utility access, safety, and environmental control.

 

Strong emphasis has been placed on environmental sustainability, with features such as a zero-discharge water strategy, a dry stack tailings facility, stormwater control systems, and reinjection wells. Emergency services, communications, and workforce facilities are also well-planned, aligning with the best practices in mine development.

 

The total Life-of-Mine capital cost is estimated at $ 417 million, comprising:

 

·Pre-production capital of $ 263.6 million (23-month period),

 

·Sustaining capital of $ 136.2 million (over 17 years), and

 

·Closure capital of $ 17.2 million, mainly in Year 14.

 

The cost estimate is a Class 5 estimate (-30% / +50%) with a 12% contingency, excluding working capital, VAT, escalation, and financing. It is based on budgetary quotes, benchmarks from Latin American projects, and internal cost databases.

 

Operating costs were derived using first principles and local benchmarks. Processing, site services, and general & administrative (G&A) costs were carefully broken down and include labor, power, consumables, and maintenance.

 

22.3.1Risks

 

The most significant potential risks associated with the Project consists of the hot water management that will be encountered in the mine dewatering effort and socio-political resistance

 

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to the development of the planned mine in Guatemala. The latter is a common risk to most mining projects and can be mitigated, at least to some degree, with adequate planning and proactive management. The risk associated with water management is not entirely unknown due to the presence of existing dewatering wells and continued dewatering, treatment and discharge of underground water.

 

It is important to note that the current mine plan is based on a resource model composed exclusively of Indicated and Inferred Resources, and Inferred Resources are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves. As such, there is a significant degree of uncertainty associated with the tonnages and grades used in the sequencing.

 

The cost of grid power is based on a market survey and not an actual power supply agreement. A higher power cost would result in increased operating costs.

 

Although the local community is favorable to the development of Era Dorada as an underground mine, there is a potential risk of socio-political opposition to mine development which could adversely impact the project development schedule.

 

The ability to achieve the estimated CAPEX and OPEX costs are important elements of Project success. If OPEX increases then the NSR cut-off would increase and, all else being equal, the size of the mineable resource would reduce yielding fewer mineable tonnes.

 

22.4Environmental Studies, Permitting, and Plans, Negotiations, or Agreements with Local Individuals or Groups

 

The project is fully permitted and has an approved EIA in place for underground mine; however, permit amendments are required for some of the proposed modifications, including increased processing rate and reinjection of mine water. New EIAs and permits are also needed for the power line. Potential delays in approval of permit amendments and/or new permits could result in increased duration of the assumed project development schedule.

 

Tailings and waste rock are assumed to be Non-Acid- Generating (NAG) based on test work completed to date and the limited exposure time at surface for waste rock. Additional test work is required prior to detailed engineering to confirm this assumption. If classification is changed to Potentially-Acid Generating (PAG), the design will need to be updated accordingly.

 

Although the local community is favorable to the development of Era Dorada as an underground mine, there is a potential risk of socio-political opposition to mine development which could adversely impact the project development schedule.

 

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23RECOMMENDATIONS

 

23.1Exploration, Geology & Resources

 

Additional drilling will increase resources and improve understanding and modeling of lithological units. Definition drilling ahead of blasting will improve the definition of grade boundaries between high-grade veins and low-grade disseminated mineralized material and help minimize unplanned dilution.

 

A review of mineral resource classification and grade distributions is prudent to ensure accuracy and certainty.

 

For geotechnical purposes, it is available to characterize and model the geotechnical parameters as domains and placement into the estimation block model.

 

A comprehensive brownfields exploration program along trend of the main deposit is recommended to explore for additional gold and silver resources that could potentially extend the project’s life.

 

23.2Mineral Processing and Metallurgical Testing and Processing and Recovery Methods

 

Based on the metallurgical test work program and the selected process route, it is recommended:

 

·To evaluate the best configuration for the comminution circuit – Primary crushing followed by a Single Stage SAG milling (SSSAG), or multi-staged crushing followed by a two-staged ball milling

 

·To perform a trade-off study comparing the CIL circuit with CIP and CIP pumpcell, all based on costs, inventory carbon and other parameters.

 

·To increase the residence time for cyanide destruction circuit to ensure the minimum residence time for coping with situations of reduced tank operation.

 

23.3Mining Methods, Infrastructure, Capital and Operating Costs

 

It is recommended for the mining methods and mining planning:

 

·Optimize the project scheduling to prioritize higher-grade zones during the initial years of operation, thereby enhancing early revenue generation.

 

·Evaluate alternative production scenarios involving variable feed rates throughout the LoM to improve project flexibility and economic performance.

 

·Conduct a PFS or FS study for Mineral Reserve certification considering potential variations in mining methods and/or stope geometry to identify opportunities for improved resource recovery and economic efficiency.

 

·Implementation of power generation in the cooling of the mine water.

 

·Mining Study detailing mining dilution for both mining methods.

 

·Detailed groundwater and dewatering control along LoM.

 

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·Develop a detailed mining operating plan that respects all the mining activities, accounting project restrictions, equipment productivities and limitations.

 

·Complete detailed engineering for site infrastructure, ensuring optimization of costs, constructability, and operational integration.

 

·Submit permitting documentation and ensure all facilities are compliant with local, national, and international environmental standards and regulations.

 

·Detailed geochemical testing for waste rock and tailings to confirm long-term environmental stability and support facility design.

 

·Maintain proactive communication with local communities and stakeholders to support social license and minimize construction-related disruptions.

 

·Implement a robust risk mitigation plan for infrastructure development, including contingency planning for stormwater events, equipment delays, and logistics challenges.

 

·Undertake a comprehensive technical and economic evaluation of the dewatering system to identify opportunities for cost reduction and efficiency improvements.

 

·Refine cost estimates to Class 5 level or higher, incorporating detailed engineering, contractor bids, and updated procurement quotes to improve accuracy and reduce contingency requirements.

 

·Evaluate project economics under different gold price scenarios, inflation rates, and cost escalations to test project resilience and identify key cost drivers.

 

·Use the current capital structure and cost estimates to support investment discussions, including potential financing, offtake agreements, or joint venture opportunities.

 

·Incorporate local tax regimes, VAT recoverability, depreciation schedules, and financing structures to derive a complete economic picture for stakeholders.

 

·Ensure that projected expenditures for G&A, environmental compliance, and social responsibility are transparently communicated and aligned with local expectations.

 

·Evaluate alternative technologies, energy-saving strategies, and hydrological modeling to minimize the operational impact of dewatering on OPEX.

 

·Implement a continuous monitoring strategy for gold price fluctuations, with regular updates to the economic model to assess impacts on Net Present Value (NPV), Internal Rate of Return (IRR), and payback period.

 

·Perform updated sensitivity analyses at key decision points to evaluate the project's resilience under various pricing scenarios.

 

23.4Environmental Studies, Permitting, and Plans, Negotiations, or Agreements with Local Individuals or Groups

 

Continuous efforts in obtaining the environmental permit amendments for groundwater injection and new EIA/permits for the power line, while advancing key activities that will reduce and de-risk the project execution schedule.

 

Costs for additional geochemical testing should be included in the budget and should be carried out prior to detailed engineering.

 

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Continue to monitor and update stakeholder engagements through the site Community Relations team. The development of close relationships with the local communities, landowners and government along with implementation of the Environmental Management Plan (EMP) and Social Management Plan (SMP) is required.

 

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24REFERENCES

 

Albinson, T. 2019. Petrographic and fluid inclusion study of samples from the Cerro Blanco project, Guatemala.

 

BaseMet Labs. 2018. Process optimization and tailings generation – Cerro Blanco Project – BL0246. 26 set. 2018.

 

Bluestone. 2019. Feasibility Study NI 43-101 Technical Report Cerro Blanco Project Guatemala. JDS Energy & Mining Inc. Data-base: 29 jan. 2019.

 

Canadian Institute of Mining, Metallurgy and Petroleum (CIM). 2014. CIM Definition Standards on Mineral Resources and Reserves. Adopted by CIM Council May 2014.

 

Canadian Institute of Mining, Metallurgy and Petroleum (CIM). 2019. CIM Estimation of Mineral Resources and Mineral Reserves Best Practice Guidelines. Adopted by CIM Council November 2019.

 

Canadian Institute of Mining, Metallurgy and Petroleum (CIM). 2019. CIM Mineral Exploration Best Practice Guidelines. Adopted by CIM Council November 2018.

 

Coplen, T. B.; Herczeg, A. L.; Barnes, C. 1999. Isotope engineering — using stable isotopes of the water molecule to solve practical problems. In: Cook, P. G.; Herczeg, A. L. (Eds.). Environmental tracers in subsurface hydrology. Boston: Springer, p. [sem paginação indicada].

 

Corbett, G. J. 1998. Epithermal gold for explorationists. AIG Journal, Paper 2002-1, fev. 2002.

 

Corporación Ambiental, S. A. 2007. Proyecto Minero Cerro Blanco: Estudio de Evaluación de Impacto Ambiental – EIA. Guatemala, jun. 2007.

 

Donnelly, T. H.; Shergold, J. H.; Southgate, P. N.; Barnes, C. J. 1990. Events leading to global phosphogenesis around the Proterozoic Cambrian boundary. Geological Society of America Bulletin, 52: 273–287.

 

Entre Mares de Guatemala, S. A.; Corporación Ambiental. 2007. Estudio de Avaliação de Impacto Ambiental (EIA) – Proyecto Minero Cerro Blanco, Municipio de Asunción Mita, Departamento de Jutiapa. Guatemala, jun. 2007.

 

Geomega Inc. 2015. Cerro Blanco materials characterization report. 9 p.

 

Golder. 2012. Initial waste rock characterization results for waste rock geochemical characterization for Cerro Blanco. 10 p. 8 jun. 2012.

 

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Hedenquist, J. W.; Arribas, A. Jr.; Gonzalez-Urien, E. 2000. Exploration for epithermal gold deposits. Reviews in Economic Geology, 13: 245–277.

 

HSRC. 2017. Health, Safety and Reclamation Code for Mines in British Columbia. Prepared by the Ministry of Energy and Mines, June 2017.

 

G Mining Services. 2021. Interim Project Readiness Report Update 2.0 – Bluestone Resources. January 2021.

 

Pindell, J.; Barrett, S. 1990. Geological evolution of the Caribbean region: a plate tectonic perspective. In: Dengo, G.; Case, J. E. (Eds.). The Caribbean Region. Geological Society of America, The Geology of North America, Volume H, p. 405–432.

 

Pindell, J.; Barrett, S. 1990. Geological evolution of the Caribbean region: a plate tectonic perspective. In: Dengo, G.; Case, J. E. (Eds.). The Caribbean Region. Geological Society of America, The Geology of North America, Volume H, p. 405–432.

 

JDS Energy & Mining Inc. 2019. Feasibility Study NI 43-101 Technical Report Cerro Blanco Project Guatemala. Effective date: 29 jan. 2019.

 

Lindgren, W. 1933. Mineral deposits. 4. ed. New York: McGraw-Hill, 930 p.

 

MEND. 2009. Prediction manual for drainage chemistry from sulphidic geologic materials. MEND Report 1.20.1.

 

Mitchell, R. J. 1983. Earth structures engineering. Winchester, MA: Allen & Unwin Inc.

 

MWH. 2014. Cerro Blanco hydrogeologic data gap, post-closure hydrogeologic conditions, and groundwater management upside potential evaluation. Project No. 10504660. May 2014.

 

Pindell, J. L.; Barrett, S. F. 1990. Geological evolution of the Caribbean region: a plate-tectonic perspective. In: Dengo, G.; Case, J. E. (Eds.). The Caribbean Region. The Geology of North America, Volume H. Geological Society of America.

 

BaseMet Labs. 2018. Process optimization and tailings generation – Cerro Blanco Project – BL0246. 26 set. 2018.

 

Rhys, D. A.; Lewis, P. D.; Rowland, J. V. 2020. Structural controls on ore localization in epithermal gold-silver deposits: a mineral systems approach. Reviews in Economic Geology, 21. Society of Economic Geologists.

 

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Savinova, E. 2020. Hydrothermal alteration mineralogy, zoning and paragenesis at the low-sulfidation epithermal Cerro Blanco deposit, Guatemala. Dissertação (Mestrado em Geociências) – University of Western Australia, Perth.

 

Sillitoe, R. H.; Hedenquist, J. W. 2003. Linkages between volcanotectonic settings, ore-fluid compositions, and epithermal precious metal deposits. Society of Economic Geologists Special Publications, 10: 315–343.

 

Sillitoe, R. H. 2018. Comments on the Cerro Blanco epithermal gold-silver deposit, Guatemala. Internal company report.

 

Sillitoe, R. H. 2018. Comments on the Cerro Blanco epithermal gold-silver deposit, Guatemala. Internal company report.

 

Stantec. 2018a. DRAFT – Cerro Blanco stormwater management and water balance report. December 2018.

 

Stantec. 2018b. Cerro Blanco numerical groundwater modeling report. 30 nov. 2018.

 

Stantec. 2018c. Cerro Blanco dewatering and water disposal report. 30 nov. 2018.

 

Stantec. 2018e. Cerro Blanco tailings geochemistry. Prepared for Bluestone Resources, Inc. by Stantec.

 

Stantec. 2018f. Cerro Blanco tailings geochemistry update. Prepared for Bluestone Resources, Inc. by Stantec.

 

Stantec. 2018f. Memo: Cerro Blanco WTP – cost estimate for treatment of process plant effluent for mercury and copper. 5 out. 2018.

 

Water Management Consultants (WMC). 2006. Cerro Blanco project interim feasibility report: hydrology and geochemistry. 65 p.

 

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25RELIANCE ON INFORMATION PROVIDED BY THE REGISTRANT

 

This TRS has been prepared by GE21 for Aura. The information, conclusions, opinions, and estimates contained herein are based on:

 

·Information available to GE21 at the time of preparation of this TRS.

 

·Assumptions, conditions, and qualifications as set forth in this TRS.

 

As part of the preparation of this Technical Report Summary (TRS) for the Project, GE21 has relied on information provided by Aura concerning legal matters, including land tenure, mineral rights, surface rights, environmental permitting, and regulatory compliance in Guatemala.

 

GE21 has relied on Aura for guidance on all relevant permits, applicable taxes, royalties, and other government levies or interests.

 

The Qualified Persons have taken all appropriate steps, in their professional opinion, to ensure that the above information from Aura is sound.

 

Except as provided by applicable laws, any use of this TRS by any third party is at that party’s sole risk.

 

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Aura Minerals Inc. | Era Dorada Gold Project

SK-1300 Technical Report Summary – Initial Assessment

June, 2025