Integra Resources : MINE DEVELOPMENT ASSOCIATES – Form 6-K | MarketScreener

MINE DEVELOPMENT ASSOCIATES

A Division of RESPEC

Contents

1.0SUMMARY 1
1.1Property Description and Ownership 1
1.2Exploration and Mining History 2
1.3Geology and Mineralization 4
1.4Drilling, Database and Data Verification 4
1.5Metallurgical Testing 5
1.6Mineral Resources 6
1.7Mineral Reserves 9
1.8Mining Methods 11
1.9Processing and Recovery Methods 11
1.10Capital and Operating Costs 12
1.11Economic Analysis 14
1.12Conclusions and Recommendations 16
1.12.1Opportunities and Risks 17
1.12.2Recommended Work Program 17
2.0INTRODUCTION AND TERMS OF REFERENCE 19
2.1Project Scope and Terms of Reference 19
2.2Frequently Used Acronyms, Abbreviations, Definitions, and Units of Measure 21
3.0RELIANCE ON OTHER EXPERTS 24
4.0PROPERTY DESCRIPTION AND LOCATION 25
4.1Location 25
4.2Land Area 26
4.3Agreements and Encumbrances 28
4.4Environmental Liabilities and Permitting 30
5.0ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE, AND PHYSIOGRAPHY 32
5.1Access to Property 32
5.2Physiography 32
5.3Climate 33
5.4Local Resources and Infrastructure 33


775-856-5700

210 South Rock Blvd.

Reno, Nevada 89502

FAX: 775-856-6053

6.0HISTORY 34
6.1Carson Mining District Discovery and Early Mining: 1863 – 1942 34
6.2Historical Exploration Since the 1960s 36
6.3Modern Historical Mining: 1977 through 1998 38
6.4Historical Resource and Reserve Estimations 40
7.0GEOLOGIC SETTING AND MINERALIZATION 43
7.1Regional Geologic Setting 43
7.2Owyhee Mountains and District Geology 44
7.3DeLamar Project Area Geology 46
7.3.1DeLamar Area 46
7.3.2Florida Mountain Area 50
7.4Mineralization 53
7.4.1District Mineralization 53
7.5DeLamar Project Mineralization 54
7.5.1DeLamar Area 54
7.5.1.1Milestone Prospect 58
7.5.2Florida Mountain Area 58
8.0DEPOSIT TYPE 61
9.0EXPLORATION 63
9.1Topographic and Geophysical Surveys 63
9.1.12019 Airborne Magnetic Survey 64
9.1.22020 Induced Polarization and Resistivity Surveys 65
9.2Rock and Soil Geochemical Sampling 65
9.3Geologic Mapping 2020 – 2021 65
9.4Database Development and Checking 65
9.5Cross-Sectional Geologic Model 66
10.0DRILLING 67
10.1Summary 67
10.2Historical Drilling – DeLamar Area 68
10.2.1Continental 1966 68
10.2.2Earth Resources 1969 – 1970 68
10.2.3Sidney Mining 1972 71
10.2.4Earth Resources ~1970 – 1983 71
10.2.5NERCO 1985 – 1992 71
10.2.6Kinross 1993 – 1998 71
10.3Historical Drilling – Florida Mountain Area 72
10.3.1Earth Resources 1972 – 1976 72
10.3.2ASARCO 1977 72
10.3.3Earth Resources 1980 72
10.3.4NERCO 1985 – 1990 72
10.3.5Kinross 1995 – 1997 73
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Integra Resources Corp.

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10.4Integra Drilling 2018 -2020 73
10.4.1DeLamar Area Drilling 2018 – 2020 74
10.4.2Florida Mountain Area Drilling 2018 – 2020 75
10.5Drill-Hole Collar Surveys 75
10.6Down-Hole Surveys 76
10.7Sample Quality and Down-Hole Contamination 76
10.8Summary Statement 77
11.0SAMPLE PREPARATION, ANALYSIS, AND SECURITY 78
11.1Historical Sample Preparation and Security 78
11.2Integra Sample Handling and Security 78
11.3Historical Sample Analysis – Prior to Commercial Open-Pit Mining Operations 79
11.4Historical Sample Analysis – During Commercial Open-Pit Mining Operations 79
11.5Integra Sample Analysis 80
11.6Quality Assurance / Quality Control Programs 81
11.6.1Historical Operators 81
11.6.2Integra 82
11.7Summary Statement 88
12.0DATA VERIFICATION 90
12.1Drill-Hole Data Verification 90
12.1.1Collar and Down-Hole Survey Data 90
12.1.2Assay Data 91
12.1.3Integra Data Verification 96
12.2Additional Data Verification 96
12.3Site Inspection 97
12.4Independent Verification of Mineralization 98
12.5Metallurgical Data Verification 99
12.6Data Verification for Mine Engineering and Geotechnical 99
12.7Summary Statement 99
13.0MINERAL PROCESSING AND METALLURGICAL TESTING 101
13.1DeLamar Area Production 1977 – 1992 101
13.1.1Mill Production 1977 – 1992 101
13.1.2Cyanide Heap Leaching 1987 – 1990 102
13.2Historical Testing 1971 – 1989 102
13.2.1Mineralogy from Historical Metallurgical Studies 102
13.2.21970s Earth Resources – Hazen Testwork 103
13.2.3Nerco Minerals Heap Leach Study 1986 103
13.2.4Sullivan Gulch Testing for NERCO 1989 103
13.2.51980s Florida Mountain Testing for NERCO 104
13.3Integra 2018-2021 Metallurgical Testing 104
13.3.1Integra DeLamar Testing 105
13.3.1.1DeLamar Samples 106
13.3.1.2DeLamar Mineralogy 108
13.3.1.3DeLamar Heap-Leach Testing 110


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13.3.1.4Integra DeLamar Mill PEA Testing 118
13.3.1.5Integra DeLamar Mill PFS Testing 119
13.3.1.6DeLamar Comminution Testing 130
13.3.2Integra Florida Mountain Area Testing 130
13.3.2.1Florida Mountain Samples 130
13.3.2.2Florida Mountain Mineralogy 133
13.3.2.3Integra Florida Mountain Heap-Leach Testing 134
13.3.2.4Integra Florida Mountain Mill PEA Testing 140
13.3.2.5Florida Mountain Mill PFS Testing 143
13.3.2.6Florida Mountain Comminution Testing 150
13.4Recovery Models and Reagents 150
13.4.1Oxide and Mixed Materials Heap-leach Recovery and Reagent Estimates 150
13.4.2DeLamar Non-Oxide Mill Recovery and Reagent Estimates 151
13.4.3Florida Mountain Non-Oxide Mill Recovery and Reagent Estimates 153
13.5Summary Statement 154
14.0MINERAL RESOURCE ESTIMATES 156
14.1Introduction 156
14.2DeLamar Project Data 158
14.2.1Drill-Hole Data 159
14.2.2Topography 159
14.2.3Modeling of Historical Underground Workings 160
14.3Geological Modeling 160
14.4Deposit Geology Pertinent to Resource Modeling 160
14.5Water Table 161
14.6Oxidation Modeling 161
14.7Density Modeling 162
14.8DeLamar Area Gold and Silver Modeling 163
14.8.1Mineral Domains 163
14.8.2Assay Coding, Capping, and Compositing 170
14.8.3Block Model Coding 172
14.8.4Grade Interpolation 173
14.8.5Model Checks 174
14.9Florida Mountain Area Gold and Silver Modeling 174
14.9.1Mineral Domains 175
14.9.2Assay Coding, Capping, and Compositing 180
14.9.3Block Model Coding 181
14.9.4Grade Interpolation 182
14.9.5Model Checks 183
14.10DeLamar Project Mineral Resources 184
14.11Discussion of Resource Modeling 198
15.0MINERAL RESERVE ESTIMATES 200
15.1Introduction 200


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Integra Resources Corp.

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15.2Pit Optimization Parameters 202
15.2.1Economic Parameters 202
15.2.2Cutoff Grades 204
15.2.3Geotechnical Parameters 204
15.2.4Royalty Boundaries 210
15.3Pit Optimization with Pit by Pit Analysis 213
15.3.1DeLamar Pit Optimization 213
15.3.2Florida Mountain Pit Optimization 216
15.4Road and Ramp Design 219
15.5Pit Design 220
15.6Proven and Probable Reserves 229
16.0MINING METHODS 233
16.1WRSFs and Backfill Designs 233
16.2Mine Production Schedule 235
16.3Equipment Requirements 241
16.4Personnel Requirements 244
17.0RECOVERY METHODS 246
17.1Process Production Schedule 246
17.2Process Design 246
17.3Heap Leach Operation 249
17.3.1Heap-Leach Crushing Plant and Agglomeration 249
17.3.2Stacking and Heap Leaching 253
17.3.3Heap-Leach Solution Pond Operation 254
17.3.4Heap-Leach Production Forecasting 254
17.4Milling Operations 256
17.4.1Comminution 256
17.4.2Flotation and Regrind 256
17.4.3Concentrate Leaching 259
17.5Merrill-Crowe Plant and Refinery 260
17.6Reagents 261
17.7Tailing Storage Facilities (TSF) and Water Consumption 262
17.7.1Tailing Storage Facilities and Water Reclamation 262
17.8Water Consumption 262
17.9Blowers and Compressors 263
17.10Power Consumption 263
17.11Control Systems 264
17.12Assay and Metallurgical Laboratories 264
17.13Alternative Processing Options 264
17.13.1High-Grade Heap Leaching Ore Processing 264
17.13.2Milling Options for DeLamar Non-oxide Ore 267
17.13.3Non-oxide Ore Gravity Concentration 268
17.13.4Cleaner Flotation Stages 268
17.13.5Process Personnel and Staffing 268


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Integra Resources Corp.

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18.0PROJECT INFRASTRUCTURE 269
18.1General Arrangement and Site Access 269
18.2Heap-Leach Pad Construction 272
18.3Slaughterhouse Gulch Tailing Storage Facility 274
18.4Concentrate Leach Tailing Storage Facility 277
18.5Renewable Energy Systems 278
18.5.1Power Generation and Distribution 278
18.5.2Power Pricing 279
18.6Railveyor Haulage System 280
18.7Project Buildings 281
18.7.1Crushing Facilities 282
18.7.2Mill and Flotation Buildings 282
18.7.3Security Building at Access Gate 282
18.7.4Administration And Changing Building 283
18.7.5Truck Shop Building 283
18.7.6Truck Wash Building 283
18.7.7Merrill-Crowe Plant and Refinery Buildings 283
18.7.8Laboratory Building 283
18.8Water Management Systems 284
18.8.1Fresh Water 284
18.8.2Water Treatment Plant 286
18.9Mine Site Personnel 286
19.0MARKET STUDIES AND CONTRACTS 287
19.1Metal Pricing 287
20.0ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR COMMUNITY IMPACT 288
20.1Environmental Studies and Permitting 288
20.1.1Environmental Baseline Studies 289
20.1.2Federal Permitting 291
20.1.2.1Bureau of Land Management Plan of Operations 291
20.1.2.2U.S. Army Corps of Engineers Authorization, Section 404 of Clean Water Act 291
20.1.2.3Environmental Impact Statement 291
20.1.3Idaho Permitting 292
20.1.3.1Idaho Pollutant Discharge Elimination System Permit (“IPDES”) 292
20.1.3.2Other Major State Authorizations, Licenses, and Permits 292
20.1.4Local County Requirements 294
20.1.5Idaho Joint Review Process 294
20.2EIS/Permitting Timelines and Costs 294
20.2.1Permitting Timelines 294
20.2.2Most Likely Case EIS Cost Summary 296
20.2.3Integra Permitting Management Strategy 296
20.3Social and Community 297
20.4Dark Skies Compliant Lighting 297


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20.5Waste Characterization 297
20.6Closure and Reclamation Strategy 298
21.0CAPITAL AND OPERATING COSTS 299
21.1Mining Capital 300
21.1.1Primary Equipment 301
21.1.2Railveyor 301
21.1.3Support Equipment 302
21.1.4Blasting Equipment 302
21.1.5Mine Maintenance Capital 302
21.1.6Other Capital 302
21.1.7Mine Preproduction Costs 303
21.2Process Capital 303
21.2.1Freight 305
21.2.2Construction Support 305
21.2.3EPCM 305
21.2.4Vendor Support 306
21.2.5Spare Parts 306
21.2.6Heap-Leach Pad Capital 306
21.2.7Tailing Impoundments 306
21.3Owner and Infrastructure Capital Costs 306
21.3.1Light Vehicles 307
21.3.2Preproduction Owner Costs 309
21.4Reclamation Costs and Salvage Value 309
21.5Mine Operating Costs 309
21.5.1Mine General Services 310
21.5.2Mine Maintenance 311
21.5.3Drilling 311
21.5.4Blasting 312
21.5.5Loading 313
21.5.6Hauling 313
21.5.7Mine Support 314
21.6Process Operating Cost Summary 314
21.6.1Power 315
21.6.2Consumable Items 316
21.6.3Maintenance 316
21.6.4Supplies and Services 317
21.6.5Application of Operating Costs 317
21.7G&A Costs 319
22.0ECONOMIC ANALYSIS 321
22.1Mining Physicals 322
22.2Pre-Tax Cash Flow 324
22.3Tax Considerations & After-Tax Cash Flow 327
22.4Sensitivity Analyses 329


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Integra Resources Corp.

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23.0ADJACENT PROPERTIES 331
24.0OTHER RELEVANT DATA AND INFORMATION 332
25.0INTERPRETATION AND CONCLUSIONS 333
25.1DeLamar Project Opportunities 334
25.1.1Heap Leach Stage 1 334
25.1.2Exploration for Underground Development 334
25.1.3Other Opportunities 335
25.1.4Environment, Social, and Governance (“ESG”) 336
25.1.4.1Equipment and Vehicle Electrification 336
25.1.4.2Ancillary Waste (Office, Truck Shop, Etc.) 337
25.1.5Geothechnical Optimization 337
25.2DeLamar Project Risks 337
25.2.1Operating Risks 337
25.2.2Permitting Risks and Risk Management Strategy 338
25.2.3Climate Change Risks 339
26.0RECOMMENDATIONS 340
27.0REFERENCES 342
28.0DATE AND SIGNATURE PAGE 349
29.0CERTIFICATE OF QUALIFIED PERSONS 350

Tables

Table 1.1 Pit Optimization Cost Parameters 7
Table 1.2 Pit-Optimization Metal Recoveries by Deposit and Oxidation State 7
Table 1.3 Total DeLamar Project Gold and Silver Resources 8
Table 1.4 Gold and Silver Resources of the DeLamar and Florida Mountain Areas 9
Table 1.5 DeLamar and Florida Mountain Economic Parameters 10
Table 1.6 Total Proven and Probable Reserves, DeLamar and Florida Mountain 10
Table 1.7 Capital Cost Summary 13
Table 1.8 Operating and Total Cost Summary 14
Table 1.9 Project Sensitivity to Metal Prices 15
Table 1.10Summary of Integra Estimated Costs for Recommended Program 18
Table 2.1 Qualified Persons, Dates of Most Recent Site Visits, and Report Responsibilities 20
Table 4.1 Summary of Estimated Land Holding Costs for the DeLamar Project 28
Table 4.2 Summary of Agreements and Encumbrances 29
Table 6.1 DeLamar Mine Gold and Silver Production 1977 – 1992 39
Table 6.2 Historical Resource and Reserve Estimates 41
Table 7.1 Summary of Volcanic Rock Units in the Vicinity of the DeLamar Mine 46
Table 10.1 DeLamar Project Drilling Summary 67
Table 10.2 Historical Drilling at the DeLamar and Florida Mountain Areas 68


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Table 10.3 Integra Drilling Summary 74
Table 11.1 Integra Certified Reference Materials 83
Table 13.1 Drill Hole Composite Summary, DeLamar PFS Metallurgical Testing 106
Table 13.2 PEA Column-Leach and Bottle-Roll Tests, DeLamar and Glen Silver Bulk Samples 113
Table 13.3 PFS Column-Leach Test and Bottle-Roll Test Results, DeLamar Core Composites 115
Table 13.4 Flotation Concentrate Regrind/Cyanidation Tests, DeLamar Non-Oxide Composites 129
Table 13.5 Drill Hole Composite Summary, Florida Mountain PEA and PFS Testing 132
Table 13.6 Column-Leach and Bottle-Roll Test Results, Florida Mountain PEA and PFS Testing 137
Table 13.7 Florida Mountain 2018-2019 Gravity Concentration with Flotation of Gravity Tailing 141
Table 13.8 Florida Mountain 2018-2019 Gravity, Flotation of Tailing and Regrind Leach of Flotation Concentrate 142
Table 13.9 Florida Mountain Flotation Concentrate Regrind/Cyanidation 146
Table 13.10 PFS Heap-Leach Recovery and Reagent Consumption Estimates 151
Table 13.11 DeLamar Non-Oxide Mill Recoveries 152
Table 13.12 DeLamar Non-Oxide Reagent Estimates 153
Table 13.13 Florida Mountain Non-Oxide Overall Mill Recoveries 154
Table 13.14 Florida Mountain Non-Oxide Overall Reagent Estimates 154
Table 14.1 Integra Specific Gravity Determinations from DeLamar Deposit Drill Core 163
Table 14.2 Integra Specific Gravity Determinations from Florida Mountain Deposit Drill Core 163
Table 14.3 Approximate Grade Ranges of DeLamar Area Gold and Silver Domains 164
Table 14.4 DeLamar Area Gold and Silver Assay Caps by Domain 170
Table 14.5 Descriptive Statistics of DeLamar Area Coded Gold Assays 170
Table 14.6 Descriptive Statistics of DeLamar Area Coded Silver Assays 171
Table 14.7 Descriptive Statistics of DeLamar Area Gold Composites 171
Table 14.8 Descriptive Statistics of DeLamar Area Silver Composites 171
Table 14.9 Summary of DeLamar Area Grade Estimation Parameters 173
Table 14.10Approximate Grade Ranges of Florida Mountain Area Gold and Silver Domains 175
Table 14.11 Florida Mountain Area Gold and Silver Assay Caps by Domain 180
Table 14.12 Descriptive Statistics of Florida Mountain Area Coded Gold Assays 180
Table 14.13 Descriptive Statistics of Florida Mountain Area Coded Silver Assays 181
Table 14.14 Descriptive Statistics of Florida Mountain Area Gold Composites 181
Table 14.15 Descriptive Statistics of Florida Mountain Area Silver Composites 181
Table 14.16 Summary of Florida Mountain Area Estimation Parameters 183
Table 14.17 Pit Optimization Cost Parameters 184
Table 14.18 Pit-Optimization Metal Recoveries 185
Table 14.19 Gold-Equivalency Factors Applied to Silver Grades 186
Table 14.20 Total DeLamar Project Gold and Silver Resources 186
Table 14.21 Resource Classification Parameters 187
Table 14.22 Gold and Silver Resources of the DeLamar and Florida Mountain Areas 188
Table 14.23 Total Project In-Pit Oxide and Mixed Materials at Various Cutoffs 197
Table 14.24 Total Project In-Pit Non-Oxide Materials at Various Cutoffs 198
Table 15.1 DeLamar and Florida Mountain Economic Parameters 203
Table 15.2 DeLamar and Florida Mountain Recoveries 203
Table 15.3PFS Pit Slope Design Recommendations – DeLamar Main, Sullivan Gulch, and Milestone 209


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Table 15.4 PFS Pit Slope Design Recommendations – Florida Mountain 209
Table 15.5 PFS Slope Design Recommendations for Soil Materials 210
Table 15.6 Royalty Zones Ownership 210
Table 15.7 DeLamar Pit Optimization Results 214
Table 15.8 DeLamar Pit by Pit Results 215
Table 15.9 Florida Mountain Pit Optimization Results 217
Table 15.10 Florida Mountain Pit by Pit Results 218
Table 15.11 Total Proven and Probable Reserves, DeLamar and Florida Mountain 229
Table 15.12 Florida Mountain Reserves by Pit Phase 230
Table 15.13 DeLamar Reserves by Pit Phase 230
Table 15.14 Florida Mountain Proven and Probable Reserves 231
Table 15.15 DeLamar Proven and Probable Reserves 231
Table 15.16 Proven and Probable Reserves by Process Type 232
Table 16.1 Waste Rock Containment Requirements (With Swell) 233
Table 16.2 WRSF and Backfill Design Capacities 234
Table 16.3 Florida Mountain Mine Production Schedule 236
Table 16.4 DeLamar Mine Production Schedule 237
Table 16.5 Total PFS Mine Production Schedule 238
Table 16.6 PFS Process Production Schedule 240
Table 16.7 Leach Ore Stockpile Balance 241
Table 16.8 Mill Ore Stockpile Balance 241
Table 16.9 PFS Yearly Mine Equipment Requirements 242
Table 16.10 Schedule Efficiency 242
Table 16.11 PFS Mining Personnel Requirements 245
Table 17.1 Heap Leach and Mill Feed Schedules 248
Table 17.2 Heap Leach Major Design Criteria 249
Table 17.3 List of Main Mechanical Equipment for the Heap-Leach Crushing Plant 252
Table 17.4 List of Conveying and Stacking Equipment for the Heap-Leach Pad 253
Table 17.5 Heap-leach Gold and Silver Production Forecasts 255
Table 17.6 Typical Achievable Variances Between Modeled and Actual Metal Productions 256
Table 17.7 Main Design Criteria for Concentrator and Cyanidation Plants 257
Table 17.8 List of Main Mill Equipment 259
Table 17.9 Main Process Reagents and Consumables 261
Table 17.10 List of Blowers and Compressors for Supply Plant and Instrument Air 263
Table 17.11 Summary of Connected Power for Heap-leach and Mill Operations 263
Table 17.12 Processing Options for High-Grade Heap-Leach Ores 266
Table 17.13 Milling Options 267
Table 18.1 Power Consumption and Production Details 279
Table 18.2 Key parameters for Railveyor System 281
Table 18.3 Mine, Process and Administrative Personnel 286
Table 21.1 Capital Cost Summary 299
Table 21.2 Operating and Total Cost Summary 300
Table 21.3 Mining Capital Cost by Year 301
Table 21.4 Stage 1 (Phase 1) Capital Costs, Oxide & Mixed Ore Heap Leach/Merrill Crowe Plant 304
Table 21.5 Stage 1 (Phase 2) Capital Costs, Oxide & Mixed Ore Heap Leach/Merrill Crowe Plant 304


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Integra Resources Corp.

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Table 21.6 Stage 2 Capital Costs – Non-Oxide Ore Grinding, Flotation, Leach Plant 305
Table 21.7 Infrastructure & Owners Capital 307
Table 21.8 Light Vehicle Cost Estimate 308
Table 21.9 Yearly Mine Operating Cost Estimate 310
Table 21.10 Mine General Services, Engineering and Geology Costs 311
Table 21.11 Yearly Mine Maintenance Costs 311
Table 21.12 Yearly Drilling Costs 312
Table 21.13 Yearly Blasting Costs 312
Table 21.14 Yearly Loading Costs 313
Table 21.15 Yearly Haulage Costs 314
Table 21.16 Yearly Mine Support Costs 314
Table 21.17 Operating Costs Ratios for Deposit/Ore Types 315
Table 21.18 Power Summary for Heap Leach & Merrill Crowe Facility 316
Table 21.19 Power Summary for Non-Oxide Mill Facility 316
Table 21.20 Modified Process Operating Costs 318
Table 21.21 Yearly G&A Costs 320
Table 22.1 Yearly Mine & Process Physicals 323
Table 22.2 Pre-Tax Cash Flow 326
Table 22.3 Depreciation, Depletion, Taxes, and After-Tax Cash Flow 328
Table 22.4 Project Sensitivity to Metal Prices 329
Table 22.5 Operating Cost Sensitivity (After Tax) 329
Table 22.6 Capital Cost Sensitivity (After Tax) 329
Table 26.1 Integra Cost Estimate for the Recommended Work Program 341

Figures

Figure 1.1 Annual Operating After-Tax Cash Flow 15
Figure 1.2 After-Tax Sensitivity 16
Figure 4.1 Location Map, DeLamar Gold – Silver Project 25
Figure 4.2 Property Map for the DeLamar Project 27
Figure 5.1 Access Map for the DeLamar Project 32
Figure 6.1 Estimated Annual Production Value, Silver City (Carson) Mining District 1863-1942 35
Figure 6.2 Aerial View, Zones of Exploration and Mining Since 1969 within the DeLamar Area 37
Figure 6.3 Aerial View of the Florida Mountain (Stone Cabin Mine) Area 39
Figure 6.4 Photograph of the Reclaimed Florida Mountain (Stone Cabin) Mine Area 40
Figure 7.1 Shade Relief Map with Regional Setting of the Owyhee Mountains 43
Figure 7.2 Geologic Map of the Central Owyhee Mountains 44
Figure 7.3 Land Position Map Showing Mineralized Zones 47
Figure 7.4 Integra Generalized 2018 DeLamar Area Geology 48
Figure 7.5 Integra 2018 Schematic Cross-Section, DeLamar Area 48
Figure 7.6 Volcano-Tectonic Setting of the DeLamar Area 49
Figure 7.7 Geologic Map of Florida Mountain 51
Figure 7.8 Schematic Florida Mountain Cross Section (Looking Northeast) 52
Figure 7.9 Veins of the Historical De Lamar Mine, Elevation 6,240 Feet 55
Figure 7.10 Longitudinal Section of the Black Jack – Trade Dollar Mine 60


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Integra Resources Corp.

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Figure 8.1 Schematic Model of a Low-Sulfidation Epithermal Mineralizing System 61
Figure 9.1 Plan View of Resistivity from 2017 and 2018 IP/RES Surveys 63
Figure 9.2 Plan View of Chargeability from 2017 and 2018 IP/RES Surveys 64
Figure 10.1 Map of DeLamar Area Drill Holes 69
Figure 10.2 Map of Florida Mountain Area Drill Holes 70
Figure 11.1 CRM CDN-GS-P6A Gold Analyses – 2018-2019 Drill Programs 84
Figure 11.2 CRM SN74 Silver Analyses – 2018-2019 Drill Programs 84
Figure 11.3 All CRM Gold Analyses – 2020-2021 Drill Programs 85
Figure 11.4 All CRM Silver Analyses – 2020-2021 Drill Programs 85
Figure 11.5 Blank Gold Values vs. Gold Values of Previous Samples – 2018-2019 Drill Programs 86
Figure 11.6 Blank Gold Values vs. Gold Values of Previous Samples – 2020-2021 Drill Programs 87
Figure 11.7 RC Field Duplicate Gold Results Relative to Primary Sample Assays 88
Figure 12.1 Repeat Mine Lab Silver Assays Relative to Original Mine Lab Assays 92
Figure 12.2 Outside Lab Silver Assays Relative to Original Mine Lab Assays 94
Figure 12.3 Mine Lab Silver AA Analyses Relative to Mine Lab Silver Fire Assays 95
Figure 12.4 Mine Lab Gold AA Analyses Relative to Mine Lab Gold Fire Assays 95
Figure 13.1 Gold Recovery, Bottle-Roll Tests, DeLamar PFS Variability Composites 111
Figure 13.2 PEA Column Column-Leach Gold and Silver Recovery vs. Feed Size 114
Figure 13.3 Glen Silver Flotation Recoveries, PFS Composites 120
Figure 13.4 Sullivan Gulch Flotation Recoveries, PFS Composites 121
Figure 13.5 Grind-Cyanide Leach Testing Gold and Silver Recoveries, Sullivan Gulch Composites 125
Figure 13.6 Florida Mountain Bottle-Roll Test Recoveries, Variability Composites 135
Figure 13.7 Recovery vs. Regrind, Flotation Conc. Cyanidation, Florida Mountain 147
Figure 13.8 Gold and Silver Recovery Rates, Flotation Ro. Concentrate, Florida Mountain 148
Figure 14.1 Cross Section 2190 NW Showing Sommercamp and N. DeLamar Gold Domains 166
Figure 14.2 Cross Section 2010 NW Showing Sommercamp and N. DeLamar Silver Domains 167
Figure 14.3 Cross Section 2790 NW Showing Gold Domains at Glen Silver 168
Figure 14.4 Cross Section 2790 NW Showing Silver Domains at Glen Silver 169
Figure 14.5 Florida Mountain Cross Section 2830 N Showing Geology and Gold Domains 176
Figure 14.6 Florida Mountain Cross Section 2830 N Showing Geology and Silver Domains 177
Figure 14.7 Florida Mountain Cross Section 3280 N Showing Geology and Gold Domains 178
Figure 14.8 Florida Mountain Cross Section 3280 N Showing Geology and Silver Domains 179
Figure 14.9 Cross Section 2190 NW Showing Sommercamp – Regan and N. DeLamar Block-Model Gold Grades 189
Figure 14.11 Cross Section 2190 NW Showing Sommercamp – Regan and N. DeLamar Block-Model Silver Grades 190
Figure 14.12 Cross Section 2790 NW Showing Glen Silver Block-Model Gold Grades 191
Figure 14.13 Cross Section 2790 NW Showing Glen Silver Block-Model Silver Grades 192
Figure 14.14 Cross Section 2830 N Showing Florida Mountain Block-Model Gold Grades 193
Figure 14.15 Cross Section 2830 N Showing Florida Mountain Block-Model Silver Grades 194
Figure 14.16 Cross Section 3280 N Showing Florida Mountain Block-Model Gold Grades 195
Figure 14.17 Cross Section 3280 N Showing Florida Mountain Block-Model Silver Grades 196
Figure 15.1 DeLamar Main- Sullivan PFS Geotechnical Sectors on 2019 PEA Pits 207
Figure 15.2 Florida Mountain PFS Geotechnical Sectors on 2019 PEA Pit 208


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Figure 15.3 Florida Mountain Royalty Zones 211
Figure 15.4 DeLamar Royalty Zones 212
Figure 15.5 DeLamar Pit by Pit Graph 216
Figure 15.6 Florida Mountain Pit by Pit Graph 219
Figure 15.7 Florida Mountain Phase 1 Design 221
Figure 15.8 Florida Mountain Phase 2 Design 222
Figure 15.9 Florida Mountain Phase 3 and Ultimate Pit Design 223
Figure 15.10 DeLamar – Milestone Phase 1 Pit Design 224
Figure 15.11 DeLamar Main Pit, Phase 1 Design 225
Figure 15.12 DeLamar Main Phase 2 and Ultimate Pit Design 226
Figure 15.13 DeLamar – Sullivan Gulch Phase 1 Pit Design 227
Figure 15.14 DeLamar – Sullivan Gulch Phase 2 and Ultimate Pit Design 228
Figure 17.1 Simplified Flow Sheet of the DeLamar Project Heap-Leach Facility 250
Figure 17.2 General Layout of the DeLamar Project Process Facilities 251
Figure 17.3 Simplified Process Flow Diagram of the Concentrator and Cyanide Leach Plants 258
Figure 18.1 PFS General Arrangement Drawing 271
Figure 18.2 Heap-Leach Pad General Arrangement 273
Figure 18.3 Heap-leach Ore Stacking Plan Geometry 274
Figure 18.4 Schematic Cross Section of Tailing Impoundment Embankment 276
Figure 18.5 Embankment Components Cross Section 277
Figure 18.6 Water Systems Flow Sheet 285
Figure 22.1 Annual Operating After-Tax Cash Flow 321
Figure 22.2 Gold Equivalent Production Profile by Process Method 324
Figure 22.3 Gold Equivalent Profile by Process Method 324
Figure 22.4 After-Tax Sensitivity 330

Appendices

Appendix A Listing of Unpatented and Patented Claims and Leased Land
Appendix B Metallurgical Test Results

Frontispiece: view looking east from above the DeLamar pit area towards Florida Mountain.

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MINE DEVELOPMENT ASSOCIATES

A Division of RESPEC

1.0SUMMARY

Mine Development Associates (“MDA”), a division of RESPEC has supervised the preparation of this technical report and Pre-Feasibility Study (“PFS”) of the DeLamar gold – silver project, located in Owyhee County, Idaho, at the request of Integra Resources Corp. (“Integra”), a Canadian company listed on the TSX Venture Exchange (TSX.V:ITR) and the NYSE American Exchange (NYSE:ITRG). The DeLamar project encompasses the DeLamar and Florida Mountain deposit areas. Both deposit areas have been subject to historical underground mining in the late 1800s and early 1900s, as well as late 20th century open-pit mining. The most recent open-pit mining, which ceased in 1998, was conducted by Kinross Gold Corporation (“Kinross”).

This report has been prepared under the supervision of Thomas L. Dyer, P.E. and Senior Engineer for MDA, Michael M. Gustin, C.P.G. and Senior Geologist for MDA, Steven I. Weiss, C.P.G. and Senior Associate Geologist for MDA, Jack McPartland, Registered Member MMSA and Senior Metallurgist with McClelland Laboratories, Inc., John Welsh, P.E., of Welsh Hagen in Reno, Nevada, Matthew Sletten, P.E. and Benjamin Bermudez, P.E. of M3 Engineering in Tucson, Arizona, Art Ibrado, P.E., of Fort Lowell Consulting in Tucson, Arizona, Jay Nopola, P.E, of RESPEC in Rapid City, South Dakota, Michael Botz, P.E., of Elbow Creek Engineering in Billings, Montana, and John F. Gardner, P.E. of Warm Springs Consulting in Boise, Idaho, in accordance with the disclosure and reporting requirements set forth in the Canadian Securities Administrators’ National Instrument 43-101 (“NI 43-101”), Companion Policy 43-101CP, and Form 43-101F1, as amended. Mr. Dyer, Mr. Gustin, Mr. Weiss, Mr. McPartland, Mr. Welsh, Mr. Sletten, Mr. Bermudez, Mr. Ibrado, Mr. Botz, Mr. Nopola, and Mr. Gardner are Qualified Persons under NI 43-101 and have no affiliation with Integra, their subsidiaries, or Kinross except that of independent consultant/client relationships.

The effective date of this technical report is January 24, 2022.

1.1Property Description and Ownership

The DeLamar project includes of 790 unpatented lode, placer, and millsite claims, and 16 tax parcels comprised of patented mining claims, as well as certain leasehold and easement interests, that cover approximately 8,673 hectares (21,431 acres) in southwestern Idaho, about 80 kilometers (50 miles) southwest of Boise. The property is approximately centered at 43°00′48″N, 116°47′35″W, within portions of the historical Carson (Silver City) mining district, and it includes the formerly producing DeLamar mine last operated by Kinross. The total annual land-holding costs are estimated to be $473,244. All mineral titles and permits are held by the DeLamar Mining Company (“DMC”), an indirect, 100% wholly owned subsidiary of Integra that was acquired from Kinross through a Stock Purchase Agreement in 2017.

775-856-5700

210 South Rock Blvd.

Reno, Nevada 89502

FAX: 775-856-6053

Technical Report and Preliminary Feasibility Study, DeLamar – Florida Mountain Project

Integra Resources Corp.

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A total of 284 of the unpatented claims were acquired from Kinross, 101 of which are subject to a 2.0% net smelter returns royalty (“NSR”) payable to a predecessor owner. This royalty is not applicable to the current project resources and reserves.

There are also eight lease agreements covering 33 patented claims and five unpatented claims that require NSR payments ranging from 2.0% to 5.0%. One of these leases covers a small portion of the DeLamar area resources and one covers a small portion of the Florida Mountain area resources and reserves, with 5.0% and 2.5% NSRs applicable to maximums of $50,000 and $650,000 in royalty payments, respectively.

The property includes 1,561 hectares (3,857.2 acres) under seven leases from the State of Idaho, which are subject to a 5.0% net smelter returns production royalty plus annual payments of $27,282. The State of Idaho leases include very small portions of both the DeLamar and Florida Mountain resources and reserves.

Kinross had retained a 2.5% NSR royalty that applies to those portions of the DeLamar area claims that are unencumbered by the royalties outlined above. The royalty was subsequently sold to Maverix Metals Inc (“Maverix”). The Maverix royalty applies to more than 90% of the current DeLamar area resources and reserves, but this royalty will be reduced to 1.0% upon Maverix receiving total royalty payments of CAD$10,000,000.

DMC also owns mining claims and leased lands peripheral to the DeLamar project described above. These landholdings are not part of the DeLamar project, although some of the lands are contiguous with those of the DeLamar and Florida Mountain claims and state leases. The DMC lands peripheral to the DeLamar project have no mineral resources or mineral reserves.

The DeLamar project historical open-pit mine areas have been in closure since 2003. While a substantial amount of reclamation and closure work has been completed to date at the site, there remain ongoing water-management activities, monitoring, and reporting. A reclamation bond of $2,778,929 remains with the Idaho Department of Lands (“IDL”) and a reclamation bond of $100,000 remains with the Idaho Department of Environmental Quality. Additional reclamation bonds in the total amount of $589,144 have been placed with the U.S. Bureau of Land Management (“BLM”) for exploration activities and groundwater well installation on public lands. There are also reclamation bonds with the IDL in the total amount of $86,900 for exploration activities on IDL leased lands.

1.2Exploration and Mining History

Total production of gold and silver from the DeLamar project area is estimated to be approximately 1.3 million ounces of gold and 70 million ounces of silver from 1891 through 1998, with an additional but unknown quantity produced at the DeLamar mill in 1999. From 1876 to 1891, an estimated 1.025 million ounces of gold and 51 million ounces of silver were produced from the original De Lamar underground mine and the later DeLamar open-pit operations. At Florida Mountain, nearly 260,000 ounces of gold and 18 million ounces of silver were produced from the historical underground mines and late 1990s open-pit mining.

Mining activity began in the area of the DeLamar project when placer gold deposits were discovered in early 1863 in Jordan Creek, a short distance upstream from what later became the town site of De Lamar. During the summer of 1863, the first silver-gold lodes were discovered in quartz veins at War Eagle Mountain, to the east of Florida Mountain, resulting in the initial settlement of Silver City. Between 1876 and 1888, significant silver-gold veins were discovered and developed in the district, including underground mines at De Lamar Mountain and Florida Mountain. A total of 553,000 ounces of gold and 21.3 million ounces of silver were reportedly produced from the De Lamar and Florida Mountain underground mines from the late 1800s to early 1900s.

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The mines in the district were closed in 1914, following which very little production took place until gold and silver prices increased in the1930s. Placer gold was again recovered from Jordan Creek from 1934 to 1940, and in 1938 a 181 tonne-per-day flotation mill was constructed to process waste dumps from the De Lamar underground mine. The flotation mill reportedly operated until the end of 1942. Including Florida Mountain, the De Lamar – Silver City area is believed to have produced about 1 million ounces of gold and 25 million ounces of silver from 1863 through 1942.

During the late 1960s, the district began to undergo exploration for near-surface bulk-mineable gold-silver deposits, and in 1977 a joint venture operated by Earth Resources Corporation (“Earth Resources”) began production from an open-pit, milling and cyanide tank-leach operation at De Lamar Mountain, known as the DeLamar mine. In 1981, Earth Resources was acquired by the Mid Atlantic Petroleum Company (“MAPCO”), and in 1984 and 1985 the NERCO Mineral Company (“NERCO”) successively acquired the MAPCO interest and the entire joint venture to operate the DeLamar mine with 100% ownership. NERCO was purchased by the Kennecott Copper Corporation (“Kennecott”) in 1993. Two months later in 1993, Kennecott sold its 100% interest in the DeLamar mine and property to Kinross, and Kinross operated the mine, which expanded to the Florida Mountain area in 1994. Mining ceased in 1998, milling ceased in 1999, and mine closure activities commenced in 2003. Closure and reclamation were nearly completed by 2014, as the mill and other mine buildings were removed, and drainage and cover of the tailing facility were developed.

Total open-pit production from the DeLamar project from 1977 through 1998, including the Florida Mountain operation, is estimated at approximately 750,000 ounces of gold and 47.6 million ounces of silver, with an unknown quantity produced at the DeLamar mill in 1999. From start-up in 1977 through to the end of 1998, open-pit production in the DeLamar area totaled 625,000 ounces of gold and about 45 million ounces of silver. This production came from pits developed at the Glen Silver, Sommercamp – Regan (including North and South Wahl), and North DeLamar areas. In 1993, the DeLamar mine was operating at a mining rate of 27,216 tonnes (30,000 tons) per day, with a milling capacity of about 3,629 tonnes (4,000 tons) per day. In 1994, Kinross commenced open-pit mining at Florida Mountain while continuing production from the DeLamar mine. The ore from Florida Mountain, which was mined through 1998, was processed at the DeLamar facilities. Florida Mountain production in 1994 through 1998 totaled 124,500 ounces of gold and 2.6 million ounces of silver.

Exploration of the DeLamar project by Integra commenced in 2017. Since then, Integra has carried out geophysical and geochemical exploration programs, and on-going geologic mapping and exploration drilling programs. Drilling was on-going as of the effective date of this report.

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Integra Resources Corp.

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1.3Geology and Mineralization

The DeLamar project is situated in the Owyhee Mountains near the east margin of the mid-Miocene Columbia River – Steens flood-basalt province and the west margin of the Snake River Plain. The Owyhee Mountains comprise a major mid-Miocene eruptive center, generally composed of mid-Miocene basalt flows intruded and overlain by mid-Miocene rhyolite dikes, domes, flows and tuffs, developed on an eroded surface of Late Cretaceous granitic rocks.

The DeLamar mine area and mineralized zones are situated within an arcuate, nearly circular array of overlapping porphyritic and flow-banded rhyolite flows and domes that overlie cogenetic, precursor pyroclastic deposits erupted as local tuff rings. Integra interprets the porphyritic and banded rhyolite flows and latites as composite flow domes and dikes emplaced along regional-scale northwest-trending structures. At Florida Mountain, flow-banded rhyolite flows and domes cut through and overlie a tuff breccia unit that overlies basaltic lava flows and Late Cretaceous granitic rocks.

Gold-silver mineralization occurred as two distinct but related types: (i) relatively continuous, quartz-filled fissure veins that were the focus of late 19th and early 20th century underground mining, hosted mainly in the basalt and granodiorite and to a lesser degree in the overlying felsic volcanic units; and (ii) broader, bulk-mineable zones of closely-spaced quartz veinlets and quartz-cemented hydrothermal breccia veins that are individually continuous for only a few meters/feet laterally and vertically, and of mainly less than 1.3 centimeters (0.5 inches) in width – predominantly hosted in the rhyolites and latites peripheral to and above the quartz-filled fissures. This second style of mineralization was mined in the open pits of the late 20th century DeLamar and Florida Mountain operations, hosted primarily by the felsic volcanic units.

The fissure veins mainly strike north to northwest and are filled with quartz accompanied by variable amounts of adularia, sericite or clay, ± minor calcite. Vein widths vary from a few centimeters to several meters, but the veins persist laterally and vertically for as much as several hundreds of meters. Principal silver and gold minerals are naumannite, aguilarite, argentite, ruby silver, native gold and electrum, native silver, cerargyrite, and acanthite. Variable amounts of pyrite and marcasite with very minor chalcopyrite, sphalerite, and galena occur in some veins. Gold- and silver-bearing minerals are generally very fine grained.

The gold and silver mineralization at the DeLamar project is best interpreted in the context of the volcanic-hosted, low-sulfidation type of epithermal model. Various vein textures, mineralization, alteration features, and the low contents of base metals in the district are typical of shallow low-sulfidation epithermal deposits worldwide.

1.4Drilling, Database and Data Verification

As of the effective date of this report, the resource database includes data from 2,836 holes, for a total of 337,268 meters (1,106,522 feet), that were drilled by Integra and various historical operators at the DeLamar and Florida Mountain areas. The historical drilling was completed from 1966 to 1998 and includes 2,625 holes for a total of 275,790 meters (904,821 feet) of drilling. Most of the historical drilling was done using reverse-circulation (“RC”) and conventional rotary methods; a total of 106 historical holes were drilled using diamond-core (“core”) methods for a total of 10,845 meters (35,581 feet). Approximately 74% of the historical drilling was vertical, including all historical conventional rotary holes. At DeLamar, a significant portion of the total meterage drilled historically was subsequently mined during the open-pit operations.

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Integra Resources Corp.

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Integra commenced drilling in 2018. As of the end of December 2020, Integra had drilled a total of 60 RC holes, 140 core holes, and 11 holes commenced with RC and finished with core tails, for a total of 61,478 meters (201,699 feet) in the DeLamar and Florida Mountain areas combined. All but one of the Integra holes were angled. Integra’s drilling continued through 2021 but none of the 2021 drilling is included in the resource database used to estimate the current mineral resources of this report.

The historical portions of the current resource drill-hole databases for the DeLamar and Florida Mountain deposit areas were created by MDA using original DeLamar mine digital database files, and this information was subjected to extensive verification measures by both MDA and Integra. The Integra portions of the drill-hole databases were directly created by MDA using original digital analytical certificates in the case of the assay tables and checking against original digital records in the case of the collar and down-hole deviation tables. Through these and numerous other verification procedures summarized herein, Mr. Gustin has verified that the DeLamar project data as a whole are acceptable as used in this report.

1.5Metallurgical Testing

Metallurgical testing by Integra, generally conducted at McClelland Laboratories (“McClelland”) during 2018 through 2021, has been used to select preferred processing methods and estimate recoveries for oxide, mixed and non-oxide mineralization from both the DeLamar and Florida Mountain deposits. Samples used for this testing, primarily drill hole composites from 2018 through 2020 Integra drilling, were selected to represent the various material types contained in the current resources from both the DeLamar and Florida Mountain deposits. Composites were selected to evaluate effects of area, depth, grade, oxidation, lithology, and alteration on metallurgical response.

Bottle-roll and column-leach cyanidation testing on drill core composites from both the DeLamar and Florida Mountain deposits and on bulk samples from the DeLamar deposit have shown that the oxide and mixed material types from both deposits can be processed by heap-leach cyanidation. These materials generally benefit from relatively fine crushing to maximize heap-leach recoveries and a feed size of 80% -12.7mm (0.5 inches) was selected as optimum. Expected heap-leach gold recoveries for the oxide mineralization from both deposits (DeLamar and Florida Mountain) are consistently high (70% – 89%). Heap leach gold recoveries for the mixed mineralization are expected to average 72% for Florida Mountain and to range from 45% to 63% for the DeLamar deposit. Heap leach silver recoveries from the Florida Mountain oxide and mixed materials are expected to average 49% and 47%, respectively. Expected heap-leach silver recoveries from the DeLamar material are highly variable (11% to 74%), but generally low. A significant portion of the DeLamar oxide and mixed mineralization will require agglomeration pretreatment using cement, because of elevated clay content. None of the Florida Mountain heap-leach material is expected to require agglomeration.

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Integra Resources Corp.

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Metallurgical testing (primarily flotation and agitated cyanidation) has shown that the DeLamar non-oxide materials respond well to flotation at a moderate grind size (150 microns) for recovery of gold and silver to a flotation concentrate. The resulting flotation concentrate responds well to cyanide leaching after very fine regrinding (20 microns) for recovery of contained silver. Some gold is also recovered by cyanide leaching of the reground flotation concentrate, but those recoveries generally are low. Mineralogical examination and metallurgical testing have shown that these materials contain significant amounts of gold that are locked in sulfide mineral particles, which require oxidative pretreatment of sulfide minerals for liberation of gold before high cyanidation gold recoveries can be obtained. Expected recoveries from the DeLamar non-oxide mineralization in the planned mill circuit, consisting of grinding, flotation concentrate regrinding and cyanide leach, range from 28% to 39% for gold and from 64% to 87% for silver.

Metallurgical testing has shown that the non-oxide mineralization from the Florida Mountain deposit responds well to upgrading by flotation at a moderate grind size (150 microns) and cyanidation gold and silver recoveries from the resulting concentrates can be maximized by very fine regrinding (20 microns). In contrast to the DeLamar non-oxide materials, oxidative pretreatment of contained sulfide minerals is not required to achieve high cyanidation gold recoveries from the Florida Mountain non-oxide feeds. Recoveries expected from the Florida Mountain non-oxide mineralization in the planned mill circuit vary with feed grade, but generally are high, with maximum recoveries of 87% gold and 77% silver.

1.6Mineral Resources

Mineral resources have been estimated for both the Florida Mountain and DeLamar areas of the DeLamar project. These gold and silver resources were modeled and estimated by:

  • evaluating the drill data statistically and spatially to determine natural gold and silver populations;

  • creating low-, medium-, and high-grade mineral-domain polygons for both gold and silver on sets of cross sections spaced at 30-meter (98.4-foot) intervals;

  • projecting the sectional mineral-domain polygons horizontally to the drill data within each sectional window;

  • slicing the three-dimensionally projected mineral-domain polygons along 6-meter-spaced horizontal planes at the DeLamar area and 8-meter-spaced (26.3-foot) planes at Florida Mountain and using these slices to recreate the gold and silver mineral-domain polygons on a set of level plans for each resource area;

  • coding a block model to the gold and silver mineral domains for each of the two deposit areas using the level-plan mineral-domain polygons;

  • analyzing the modeled mineralization geostatistically to aid in the establishment of estimation and classification parameters; and

  • interpolating gold and silver grades by inverse-distance to the third power into 6 x 6 x 6-meter (19.7 x 19.7 x 19.7-foot) blocks for the DeLamar area and 6 x 8 x 8-meter (19.7 x 26.3 x 26.3-foot) blocks at Florida Mountain, using the coded gold and silver mineral-domain percentages to explicitly constrain the grade estimations.

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Integra Resources Corp.

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To meet the requirement of the in-pit resources having reasonable prospects for eventual economic extraction, pit optimizations for the DeLamar and Florida Mountain deposit areas were run using the parameters summarized in Table 1.1 and Table 1.2.

Table 1.1 Pit Optimization Cost Parameters

Parameter DeLamar Florida Mtn Unit
Mining Cost $ 2.00 $ 2.00 $/tonne mined
Heap Leach
Oxide Processing $ 2.75 $ 2.75 $/tonne processed
Mixed Processing $ 3.75 $ 3.50 $/tonne processed
Incremental Haulage $ 0.20 $ 0.20 $/tonne processed
G&A $ 0.40 $ 0.40 $/tonne processed
Mill – DeLamar Area
Non-Oxide Processing $ 15.25 $ – $/tonne processed
Incremental Haulage $ 0.20 $ – $/tonne processed
G&A Cost $ 0.25 $ – $/tonne processed
Mill – Florida Mountain Area
Non-Oxide Processing $ – $ 9.00 $/tonne processed
Incremental Haulage $ 0.20 $/tonne processed
G&A Cost $ – $ 0.25 $/tonne processed
Au Price $ 1,800 $ 1,800 $/oz produced
Ag Price $ 21 $ 21 $/oz produced
Au Refining Cost $ 5.00 $ 5.00 $/oz produced
Ag Refining Cost $ 0.50 $ 0.50 $/oz produced
Royalty see Section 4.3 see Section 4.3 NSR

Table 1.2 Pit-Optimization Metal Recoveries by Deposit and Oxidation State

DeLamar Florida Mountain
Process Type Oxide Mixed Non-Oxide Oxide Mixed Non-Oxide
Heap Leach – Au 85% 80% 90% 85%
Heap Leach – Ag 45% 40% 65% 55%
Mill – Albion – Glen Silver – Au 78%
Mill – Albion – Glen Silver – Ag 78%
Mill – Albion – Non-Glen Silver – Au 87%
Mill – Albion – Non-Glen Silver – Ag 87%
Mill – Agitated Leach – Au 95%
Mill – Agitated Leach – Ag 92%


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Integra Resources Corp.

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The DeLamar project mineral resources were estimated to reflect potential open-pit extraction and processing by: crushing and heap leaching of oxide and mixed materials at both the DeLamar and Florida Mountain areas; grinding, flotation, ultra-fine regrind of concentrates, and Albion cyanide-leach processing of the reground concentrates for the non-oxide materials at the DeLamar area; and grinding, flotation, ultra-fine regrind of concentrates, and agitated cyanide-leaching of non-oxide materials at Florida Mountain. To meet the requirement of having reasonable prospects for eventual economic extraction by open-pit methods, pit optimizations for the DeLamar and Florida Mountain areas were run using the parameters summarized in Table 1.1 and Table 1.2, and the resulting pits were used to constrain the project resources.

The pit shells created using these optimization parameters were applied to constrain the DeLamar project resources. The in-pit resources were further constrained by the application of a gold-equivalent cutoff of 0.17 g/t to all model blocks lying within the optimized pits that are coded as oxide or mixed, a 0.3 g/t gold-equivalent cutoff for blocks coded as non-oxide at DeLamar, and a 0.2 g/t cutoff for blocks coded as non-oxide at Florida Mountain. Gold-equivalent grades, which were used solely for the purpose of applying the resource cutoffs, are a function of metal prices (Table 1.1) and metal recoveries, with the recoveries varying by deposit and oxidation state (Table 1.2).

The total DeLamar project resources, including both the DeLamar and Florida Mountain areas, are summarized in Table 1.3. The project mineral resources are inclusive of the mineral reserves discussed herein. Mineral resources that are not mineral reserves do not have demonstrated economic viability.

Table 1.3 Total DeLamar Project Gold and Silver Resources

1.The effective date of the mineral resources is March 1, 2021.

2.Mineral resources are reported inclusive of mineral reserves.

3.Mineral resources that are not mineral reserves do not have demonstrated economic viability.

4.Rounding may result in slight discrepancies between tonnes, grade, and contained metal content.

5.The estimate of mineral resources may be materially affected by geology, environment, permitting, legal, title, taxation, sociopolitical, marketing, or other relevant issues.

The gold and silver resources for the DeLamar and Florida Mountain areas are reported separately in Table 1.4.

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Integra Resources Corp.

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Table 1.4 Gold and Silver Resources of the DeLamar and Florida Mountain Areas

1.The effective date of the mineral resources is March 1, 2021.

2.Mineral resources are reported inclusive of mineral reserves.

3.Mineral resources that are not mineral reserves do not have demonstrated economic viability.

4.Rounding may result in slight discrepancies between tonnes, grade, and contained metal content.

5.The estimate of mineral resources may be materially affected by geology, environment, permitting, legal, title, taxation, sociopolitical, marketing, or other relevant issues.

1.7Mineral Reserves

Mr. Dyer has used Measured and Indicated mineral resources as the basis to define mineral reserves for both the DeLamar and Florida Mountain deposits. Mineral reserve definition was done by first identifying ultimate pit limits using economic parameters and pit optimization techniques. The resulting optimized pit shells were then used for guidance in pit design to allow access for equipment and personnel. Mr. Dyer then considered mining, processing, metallurgical, infrastructure, economic, marketing, legal, environmental, social, and governmental factors for defining the estimated mineral reserves.

The economic parameters and cutoff grades used in the estimation of the mineral reserves are shown in Table 1.5 The overall leaching process rate is planned to be 35,000 tonnes (38,581 tons) per day or 12,600,000 tonnes (13,889,123 tons) per year for both Florida Mountain and DeLamar oxide and mixed material. DeLamar leach processing will also include agglomeration. Initially only the oxide and mixed material will be processed, then starting in year 3, non-oxide will be processed through a plant constructed to operate at a rate of 6,000 tonnes (6,614 tons) per day or 2,160,000 tonnes (2,380,992 tons) per year.

The cutoff grades applied reflect the cost to process material along with G&A and incremental haulage costs. Note that royalties are built into the block values and are considered in determining whether to process the material. While the DeLamar non-oxide breakeven cutoff grade would be $11.44/t according to the applicable costs, a cutoff of $15.00 was assigned to enhance the project’s economic performance.

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Integra Resources Corp.

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Table 1.5 DeLamar and Florida Mountain Economic Parameters

GMV = gross metal value; COG = cutoff grade.

Total Proven and Probable reserves for the DeLamar project from all pit phases are 123,483,000 tonnes at an average grade of 0.45 g Au/t and 23.27 g Ag/t, for 1,787,000 ounces of gold and 92,403,000 ounces of silver (Table 1.6). The mineral reserves point of reference is the point where material is fed into the crusher.

Table 1.6 Total Proven and Probable Reserves, DeLamar and Florida Mountain

Notes:

(1)All estimates of Mineral Reserves have been prepared in accordance with National Instrument 43 – 101 – Standards of Disclosure for Mineral Projects (“NI 43-101”) and are included within the current Measured and Indicated mineral resources.

(2)Thomas L. Dyer, PE for MDA, a division of RESPEC, in Reno, Nevada, is a Qualified Person as defined in NI 43-101, and is responsible for reporting Proven and Probable mineral reserves for the DeLamar Project. Mr. Dyer is independent of Integra.

(3)Mineral reserves are based on prices of $1,650 per ounce Au and $21.00 per ounce Ag. The reserves were defined based on pit designs that were created to follow optimized pit shells created in Whittle. Pit designs followed pit slope recommendations provided by RESPEC.

(4)Reserves are reported using block value cutoff grades representing the cost of processing:

Florida Mountain oxide leach cutoff grade value of $3.55/t.

Florida Mountain mixed leach cutoff grade value of $4.20/t.

Florida Mountain non-oxide mill cutoff grade value of $10.35/t.

DeLamar oxide leach cutoff grade value of $3.65/t

DeLamar mixed leach cutoff grade value of $4.65/t.

DeLamar non-oxide mill cutoff grade value of $15.00/t.

(5)The mineral reserves point of reference is the point where is material is fed into the crusher.

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(6)The effective date of the mineral reserves estimate is January 24, 2022.

(7)All ounces reported herein represent troy ounces, “g Au/t” represents grams per gold tonne and “g Ag/t” represents grams per silver tonne.

(8)Columns may not sum due to rounding.

(9)The estimate of mineral reserves may be materially affected by geology, environment, permitting, legal, title, taxation, sociopolitical, marketing, or other relevant issues.

(10)Energy prices of US$2.50 per gallon of diesel and $0.065 per kWh were used.

1.8Mining Methods

The PFS presented in this report considers open-pit mining of the DeLamar and Florida Mountain gold-silver deposits. Mining will utilize 23-cubic meter (30-cubic yard) hydraulic shovels along with 13-cubic meter (16.7-cubic yard) loaders to load 136-tonne capacity haul trucks. The haul trucks will haul waste and ore out of the pit and to dumping locations. Due to the length of ore hauls, the ore will be stockpiled near the pits followed by loading into a Railveyor system which will convey the ore into a crusher. The Railveyor system will be supplemented with haul trucks on an as needed basis.

Waste material will be stored in waste-rock storage facilities (“WRSFs”) located near each of the Florida Mountain and DeLamar deposits, as well as backfilled into pits where available. The exception is the Milestone pit, from which waste material will be fully utilized for construction material for the tailing storage facility (“TSF”).

Production scheduling was completed using Geovia’s MineSched™ (version 2021) software. Proven and Probable reserves along with waste material inside pit designs previously discussed were used to schedule mine production. The production schedule considers the processing of DeLamar and Florida Mountain oxide and mixed material by crushing and heap leaching, with some of the DeLamar material requiring agglomeration prior to leaching. DeLamar and Florida Mountain non-oxide material would be processed using flotation followed by cyanide leaching of the flotation concentrate.

An autonomous Railveyor light-rail haulage system will be used to transport ore from the open pits to the crusher facility. Utilizing the Railveyor system allows the opportunity to realize cost savings compared to typical truck haulage. This system, in conjunction with the planned solar and liquid natural gas electrical microgrid will reduce the overall fuel consumption and carbon footprint of the project.

The PFS has assumed owner mining instead of the more expensive contract mining. The production schedule was used along with additional efficiency factors, performance curves, and productivity rates to develop the first-principal hours required for primary mining equipment to achieve the production schedule. Primary mining equipment includes drills, loaders, hydraulic shovels, and haul trucks. Support, blasting, and mine maintenance equipment will be required in addition to the primary mining equipment.

1.9Processing and Recovery Methods

This PFS envisions the use of two process methods for the recovery of gold and silver:

  • Lower-grade oxide and mixed materials will be processed by crushed-ore cyanide heap leaching; and

  • Non-oxide material will be processed using grinding followed by flotation, and very fine grinding of flotation concentrate for agitated cyanide leaching.

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Heap-leach and milling ores will be coming from both the Florida Mountain and DeLamar deposits. Pregnant solutions from the heap-leach operation and from the milling operation will be processed by the same Merrill-Crowe zinc cementation plant. Processing will start with heap leaching in the first two years of operation. Milling of higher-grade non-oxide ore will start in the third year of operation.

Both Florida Mountain and DeLamar oxide and mixed ore types have been shown to be amenable to heap-leach processing following crushing. Material will be crushed in three stages to a nominal size of 80% finer than (P80) 12.7-millimeter (0.5 inches), at a rate of 35,000 tonnes per day. About 45% of DeLamar ore is expected to require agglomeration.

Crushed and prepared ore will be transferred to the heap-leach pad using overland conveyors and stacked on the heap using portable or grasshopper conveyors and a radial stacking system. Pregnant leach solution will be collected at the base on the heap leach and transferred to the Merrill-Crowe processing plant for recovery of precious metals by zinc precipitation. The precipitate will be filtered, dried, and smelted to produce gold and silver doré bullion for shipment off site.

The milling process will start with primary crushing of the ore to a nominal P80 of 120 millimeter (4.72 inches), followed by grinding in a SAG mill-ball mill circuit to a P80 of 150 microns. The ball mill discharge will be pumped to hydrocyclones, with the hydrocyclone overflow advancing to flotation and the underflow returning to the ball mill.

The flotation circuit will produce a sulfide concentrate that will recover gold and silver from the ore. This flotation concentrate will be reground to a nominal P80 of 20 microns before being leached in agitated leach tanks. Pregnant solution will be separated using a CCD circuit that employs dewatering cyclones and thickeners. The pregnant solution is then sent to the Merrill-Crowe plant and gold smelting facility to produce gold and silver doré bullion.

The flotation tailing stream will be thickened and pumped to the tailing storage facility. The concentrate leach residue will be sent to cyanide destruction, then stored in a separate concentrate leach tailing storage facility.

1.10Capital and Operating Costs

Table 1.7 summarizes the estimated capital costs for the project. The life of mine (“LOM”) total capital costs are estimated as $589.5 million, including $307.6 million in preproduction capital (including working capital and reclamation bond) and $281.8 million for expansion and sustaining capital. Sustaining capital includes $30.8 million in reclamation costs. The estimated capital costs are inclusive of sales tax, engineering, procurement and construction management (“EPCM”) and contingency.

Table 1.8 shows the estimated LOM operating costs for the project. Operating costs are estimated to be $12.93 per tonne processed for the LOM. This includes mining costs, which are estimated to be $1.90 per tonne mined. The total cash cost is estimated to be $923 per ounce of gold equivalent and site level all-in sustaining costs are estimated to be $955 per ounce of gold equivalent.

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Table 1.7 Capital Cost Summary

Notes:

(1)Capital costs include contingency and EPCM costs.

(2)Mining equipment includes cost of Railveyor.

(3)Major mining equipment assumes financing by equipment vendor with 10% down; principal payments included under sustaining capital column and interest payments included in operating costs.

(4)Sustaining capital showed in this table includes expansion capital (non-oxide plant) and principal payment of mining equipment leases (see note 3 above).

(5)Working capital is returned in year 17.

(6)Cash deposit = 20% of bonding requirement. Released once reclamation is completed.

(7)Salvage value for mining equipment and plant.

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Table 1.8 Operating and Total Cost Summary

Notes:

(1)By-Product costs are shown as US dollars per gold ounces sold with silver as a credit; and

(2)Co-Product costs are shown as US dollars per gold equivalent ounce.

1.11Economic Analysis

Economic highlights of this PFS for the DeLamar mining project include:

  • Initial construction period is anticipated to be 18 months;

  • After-tax net present value (“NPV”) (5%) of $407.8 million with a 27% after-tax internal rate of return (“IRR”) using $1,700 and $21.50 per ounce gold and silver prices, respectively;

  • After-tax payback period of 3.34 years;

  • Year 1 to 8 gold equivalent average production of 163,000 ounces (average 121,000 oz Au/year and 3,312,000 oz Ag/year);

  • Year 1 to 16 gold equivalent average production of 110,000 ounces (average 71,000 oz Au/year and 3,085,000 oz Ag/year).

  • After-tax LOM cumulative cash flow of $689.3 million; and

  • Average annual after-tax free cash flow of $59.8 million during production.

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Figure 1.1 shows the annual operating after-tax cash flow.

Figure 1.1 Annual Operating After-Tax Cash Flow

Economic sensitivities of the project to changes in metal prices were evaluated based on constant gold to silver ratios as shown in Table 1.9. The after-tax sensitivity to revenues, capital, and operating costs is shown in Figure 1.2.

Table 1.9 Project Sensitivity to Metal Prices

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Figure 1.2 After-Tax Sensitivity

1.12Conclusions and Recommendations

Integra has advanced the DeLamar project to report herein the first Proven and Probable mineral reserves since mining by Kinross ceased in 1998. The PFS indicates a strong economic viability as shown with (i) after-tax NPV (5%) of $407.8 million with a 27% after-tax IRR using prices of $1,700 and $21.50 per ounce gold and silver, respectively; (ii) after-tax payback period of 3.34 years; and (iii) year 1 to 8 average production of 163,000 oz AuEq (121,000 oz Au and 3,312,000oz Ag), with year 1 to 16 average production of 110,000 oz AuEq (71,000 oz Au and 3,085,000 oz Ag). The total cash cost is estimated to be $923 per oz AuEq, with site-level all-in sustaining costs estimated to be $955 per oz AuEq.

The PFS total LOM gold production is estimated to be 1,154,000 ounces, with LOM average recovery of 72% for the heap leach and 51% for the mill. Silver production is estimated to be 50.0 million ounces, with an average LOM recovery of 37% for the heap leach and 75% for the mill.

Project-wide Measured and Indicated resources, inclusive of Proven and Probable reserves, total 200,248,000 tonnes averaging 0.40 g Au/t (2,597,000 ounces of gold) and 19.7 g Ag/t (126,968,000 ounces of silver). Inferred resources total 40,615,000 tonnes at an average grade of 0.35 g Au/t (452,000 ounces of gold) and 12.5 g Ag/t (16,358,000 ounces of silver). Total Proven and Probable reserves for the DeLamar project from all pit phases are 123,483,000 tonnes at an average grade of 0.45 g Au/t and 23.27 g Ag/t, for 1,787,000 ounces of gold and 92,403,000 ounces of silver.

Metallurgical testing has shown that oxide and mixed mineralization types from both the DeLamar and Florida Mountain deposits can be processed by heap-leach cyanidation, with no need for agglomeration pretreatment of Florida Mountain material, where production starts, but with agglomeration required for a significant portion of the DeLamar area heap-leachable mineralization. Non-oxide mineralization from the Florida Mountain and DeLamar deposits is amenable to grinding followed by flotation, flotation concentrate regrind, and agitated cyanide leaching of the reground concentrate for recovery of gold and silver. The average silver recovery from the DeLamar non-oxide material is 75%, which leads to silver making a significant contribution to the project economics.

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There is considerable exploration upside for both potential open-pit and underground mineable mineralization, the former at Florida Mountain and in Integra-held lands immediately outside of the bounds of the PFS project area. The potential for underground mineable mineralization is particularly prospective in the Florida Mountain area. The potential of both areas and target types is supported by existing Integra drill results and warrants significant additional exploration investment.

1.12.1Opportunities and Risks

There are many opportunities to improve the DeLamar project as discussed in Section 25, which in summary include (i) exploration both within the PFS project boundary and immediately outside of this boundary on other lands held by Integra that leads to additional open-pit and underground resources; (ii) additional metallurgical testing that results in improved process strategies and recoveries and/or decreased process cost; (iii) evaluation of the historical waste dumps and backfill that determines their suitability for being processed in a similar manner as the current reserves; (iv) improved understanding of the geotechnical aspects of the deposits that allows for steepening of the current slope parameters; and (v) the development of strategies for electrification of operations to decrease costs and CO2 emissions.

An additional opportunity is the possibility of developing the DeLamar project deposits through a heap-leach only operation, which would lower the LOM capital costs and lower the operating costs while maintaining a robust production profile. In this scenario, the decision to construct and initiate mill processing could be exercised at any time, providing the flexibility to respond to changing market conditions and thereby reduce project risk.

Project risks include: (i) heap-leach gold and silver recoveries from DeLamar mixed materials and mill gold recoveries from non-oxide materials are variable and may not fully achieve projected recoveries; (ii) elevated-clay material at the DeLamar area may adversely affect projected heap leach and/or mill recoveries; (iii) further geotechnical studies for leach pads may result in less favorable geotechnical parameters that could add costs and larger footprints of heap-leach pads; and (iv) the hydrogeology is not well understood at present, which could lead to higher than anticipated water-management costs.

1.12.2Recommended Work Program

Significant additional work is recommended to advance the DeLamar project. The total cost for the recommended program is approximately $20.2 million (Table 1.10), including work towards the completion of an updated PFS in 2022, work to optimize metallurgical recoveries, develop a feasibility study, and permitting studies with the goal to file a BLM Plan of Operations in mid-2023. The estimated drilling costs are all-inclusive, as they include Integra’s labor, drilling costs, access and drill-pad construction costs, assaying, etc., in addition to the contractor costs.

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Table 1.10 Summary of Integra Estimated Costs for Recommended Program

Item

Estimated Cost US$

Exploration Drilling (8,000 meters)

$4,000,000

Geological Mapping, Soil Sampling

$400,000

Metallurgical Core Drilling (5,000 meters)

$2,400,000

Condemnation RC Drilling (3,000 meters)

$1,200,000

Legacy waste-rock / storage facilities and pit backfill Drilling (1,500 meters)

$350,000

Geotechnical Core and Auger Drilling (2,500 meters)

$900,000

Groundwater RC Drilling (1,650 meters)

$650,000

Metallurgical Testwork

$1,000,000

Engineering, Design, and Geotechnical Studies

$3,100,000

Permitting

$6,200,000

Total

$20,200,000


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2.0INTRODUCTION AND TERMS OF REFERENCE

Mine Development Associates (“MDA”) a division of RESPEC has supervised the preparation of this technical report and Pre-Feasibility Study (“PFS”) of the DeLamar and Florida Mountain gold-silver project (“the DeLamar project”), located in Owyhee County, Idaho, at the request of Integra Resources Corp. (“Integra”), a Canadian company based in Vancouver, British Columbia. Integra entered into a binding stock purchase agreement dated September 18, 2017 with Kinross Gold Corporation (“Kinross”) to acquire the Kinross DeLamar Mining Company, then an indirect, wholly owned subsidiary of Kinross, and thereby acquired 100% of its DeLamar gold-silver property. Subsequent to that transaction, Integra has acquired 100% interests in significant additional lands at the adjacent Florida Mountain property, as well as other lands outside of the limits of the PFS project.

Integra is listed on the TSX Venture Exchange (TSX.V: ITR) and the NYSE American Exchange (NYSE:ITRG). This report draws from previous technical reports by Gustin and Weiss (2017), and Gustin et al., (2019a; 2019b), and has been prepared in accordance with the disclosure and reporting requirements set forth in the Canadian Securities Administrators’ National Instrument 43-101 (“NI 43-101”), Companion Policy 43-101CP, and Form 43-101F1, as amended.

2.1Project Scope and Terms of Reference

The purpose of this report is to provide a PFS and updated technical summary of the DeLamar gold-silver project. This PFS includes the first estimate of mineral reserves for the project. The estimated reserves are based on updated estimates of mineral resources, reported herein, that have an effective date of March 1, 2021 and are the current mineral resources for the project. Mineral reserves reported herein have an effective date of January 24, 2022.

The DeLamar project lies within the historical Carson (Silver City) mining district of southwestern Idaho. The most recent production from the project occurred in 1977 through 1998 by open-pit mining with both milling and minor cyanide heap-leach processing of gold-silver ores. The mine was placed on care and maintenance in 1999, and later underwent mine closure by Kinross.

In addition to the estimation of reserves and the updated DeLamar and Florida Mountain mineral resources, the scope of the work completed by the authors included a review of pertinent technical reports and data provided to the authors by Integra relative to the general setting, geology, project history, exploration and mining activities and results, drilling programs, methodologies, quality assurance, metallurgy, and interpretations. This work culminated in the estimation of mineral resources and reserves. References are cited in the text and listed in Section 20.0.

This report has been prepared under the supervision of Thomas L. Dyer, P.E., and Principal Engineer for MDA, Michael M. Gustin, C.P.G. and Senior Geologist for MDA, Jay R. Nopola, P.E. for RESPEC, Steven I. Weiss, C.P.G. and Senior Associate Geologist for MDA, Jack McPartland, Registered Member MMSA and Senior Metallurgist with McClelland Laboratories, Inc., Matthew Sletten, P.E. and Benjamin Bermudez, P.E. for M3 Engineering of Tucson, Arizona, Art S. Ibrado, P.E., of Fort Lowell Consulting PLLC., in Tucson, Arizona, John Welsh, P.E., of Welsh Hagen and Associates in Reno, Nevada, John F. Gardner, P.E., of Warm Springs Consulting in Boise, Idaho, and Michael M. Botz, P.E. for Elbow Creek Engineering Inc., in Billings, Montana. Mr. Dyer, Mr. Gustin, Mr. Weiss, Mr. Nopola, Mr. McPartland, Mr. Sletten, Mr. Bermudez, Mr. Ibrado, Mr. Welsh, Mr. Gardner and Mr. Botz are Qualified Persons under NI 43-101 and have no affiliation with Integra except that of independent consultant/client relationships.

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Table 2.1 lists the preparers of this report, all of which are qualified persons, as well as the sections of the report for which they are responsible and the date of their most recent site inspection, where applicable.

Table 2.1 Qualified Persons, Dates of Most Recent Site Visits, and Report Responsibilities

Company

Qualified Person

Professional

Designation

Date of Most Recent

Site Visit

QP Responsibilities

by Report Section

Mine Development Associates,

a division of RESPEC

Tom Dyer

P.E.

10/27/2020

1.7, 1.8, 1.10, 1.11, 1.12, 15 (except for 15.2.3), 16, 18.1, 18.5, 18.6, 18.9, 19, 21 (except for 21.2, 21.6.1 through 21.6.5), 22, 25

26

Michael Gustin

C.P.G.

10/27/2020

1.1, 1.2, 1.3, 1.4, 1.12, 2, 3, 4, 5, 6, 7, 8, 9, 10, 11, 12, 14, 20, 23, 24, 25.1, 25.2, 26, 27, 28, 29

Steven Weiss

C.P.G.

8/3/2017

1.1, 1.2, 1.3, 1.4, 4, 5, 6, 7, 8, 9, 10, 11, 20, 25.1, 26, 27, 28, 29

RESPEC

Jay R. Nopola

P.E.

9/24/2020

15.2.3

McClelland Laboratories

Jack McPartland

MMSA

1/17/2019

1.5, 13

M3 Engineering & Technology

Matthew Sletten

P.E.

10/27/2020

18.7 and 18.8

Benjamin Bermudez

P.E.

none

1.9, 17 (except 17.3.3, 17.3.4), 21.2 (except 21.2.6, 21.2.7), 21.6.1 through 21.6.5

Fort Lowell Consulting

Art Ibrado

P.E.

none

17 (except 17.3.3, 17.3.4)

Welsh Hagen Associates

John Welsh

P.E.

10/27/2020

17.3.3, 18.2 through 18.4 Parts of 21 and 25

Warm Springs Consulting

John F. Gardner

P.E.

none

18.5

Elbow Creek Engineering

Michael M. Botz

P.E.

none

17.3.2, 17.3.3, 17.3.4

Mr. Weiss visited the project site on August 1, 2, and 3, 2017, accompanied and assisted by Ms. Kim Richardson of Jordan Valley, Oregon. Ms. Richardson is a geologist who joined the DeLamar mine staff in 1980 and eventually held the positions of Senior Mine Geologist, Mine Superintendent, and Mine Manager before leaving the project in 1997. Mr. Weiss reviewed the property geology, exposures of mineralized rocks in still accessible open pits, and areas of historical exploration drilling peripheral to the open pits, as well as historical exploration data on file at the DeLamar mine-site office.

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Mr. Gustin visited the project site on October 16, 17, and 18, 2018, October 15, 2020, and October 27, 2020, accompanied by various members of the Integra technical team. Mr. Gustin received updates on the property geology, drilling results to date, and drill-targeting concepts. He also visited all relevant areas of the project, inspected mineralized drill-core intervals from various holes, and discussed details of the drilling, drill-sampling, and quality control methods and procedures with the Integra technical team. Mr. Dyer visited the project site on October 27, 2020. Mr. Nopola visited the project site on September 24, 2020.

Section 13, Mineral Processing and Metallurgical Testing, was prepared under the supervision of Mr. Jack S. McPartland, Senior Metallurgist with (“McClelland”) Laboratories, Inc., in Sparks, Nevada. Mr. McPartland visited the DeLamar project site on January 17, 2019.

Section 17 Recovery Methods was prepared under the supervision of Mr. Benjamin Bermudez, P.E., for M3 Engineering, and Mr. Art Ibrado, P.E., of Fort Lowell Consulting working with M3 Engineering. Mr. John Welsh, P.E., of Welsh Hagen Associates in Reno, Nevada, and Mr. Matthew Sletten of M3 Engineering contributed portions of Section 18 Infrastructure. Mr. Bermudez contributed to Section 21 Capital and Operating Costs. Mr. Sletten, Mr. Bermudez, Mr. Ibrado and Mr. Welsh are Qualified Persons under NI 43-101. Mr. Ibrado, Mr. Bermudez, Mr. Botz and Mr. Gardner have not visited the project site. Mr. Welsh last visited the property on October 27, 2020. Mr. Sletten visited the property on October 27, 2020.

The authors have reviewed the available data and have made judgments as to the general reliability of this information. Where deemed either inadequate or unreliable, the data were either eliminated from use or procedures were modified to account for lack of confidence in that specific information. The authors have made such independent investigations as deemed necessary in their professional judgment to be able to reasonably present the conclusions discussed herein.

This report describes the estimated mineral resources and reserves for both the DeLamar and Florida Mountain areas. To avoid potential ambiguities, the term “DeLamar project” refers to the entire project, while “DeLamar”, “DeLamar area”, or “DeLamar deposit” and “Florida Mountain”, or “Florida Mountain area” or “Florida Mountain deposit” refer to the individual areas.

The effective date of the current mineral resources that support the PFS is March 1, 2021. The effective date of the current mineral reserves is January 24, 2022 and the effective date of this technical report is also January 24, 2022.

2.2Frequently Used Acronyms, Abbreviations, Definitions, and Units of Measure

In this report, measurements are generally reported in metric units. Where information was originally reported in Imperial units, conversions have been made with the following conversion factors:

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Linear Measure

1 centimeter= 0.3937 inch

1 meter= 3.2808 feet= 1.0936 yard

1 kilometer= 0.6214 mile

Area Measure

1 hectare= 2.471 acres= 0.0039 square mile

Capacity Measure (liquid)

1 liter= 0.2642 US gallons

1 cubic meter= 264.172 US gallons

Weight

1 tonne= 1.1023 short tons= 2,205 pounds

1 kilogram= 2.205 pounds

Conversion of Imperial to Metric Grades

1 troy ounce per short ton= 34.2857 grams per metric tonne

Currency: Unless otherwise indicated, all references to dollars ($) in this report refer to currency of the United States.

Frequently used acronyms and abbreviations

AA atomic absorption spectrometry
Ag silver
Au gold
AuEq gold equivalent
cm centimeters
core diamond core-drilling method
oC degrees centigrade
CAD$ Canadian dollars
CO2e carbon dioxide equivalent
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dMTPH dry metric tonne per hour
°F degrees Fahrenheit
ft foot or feet
gpm gallons per minute
g/t grams per tonne
ha hectares
h or hr hour
ICP inductively coupled plasma analytical method
in. inch or inches
kg kilograms
km kilometers
kPa kilopascals
kt kilotonnes
ktpd metric kilotonnes per day
kV kilovolt
kWh kilowatt hour
l or L liter
lbs pounds
µm micron
m meters
m3 cubic meters
Ma million years old
mi mile or miles
mm millimeters
NPV net present value
NSR net smelter return
oz ounce
ppm parts per million
ppb parts per billion
QA/QC quality assurance and quality control
RC reverse-circulation drilling method
RQD rock-quality designation
t or mt metric tonne or tonnes
TPD metric tonnes per day
tph metric tonnes per hour
ton Imperial short ton
U.S. United States of America
MW megawatt
MWh megawatt hours
XRD x-ray diffraction
XRF x-ray fluorescence


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3.0RELIANCE ON OTHER EXPERTS

The authors are not experts in legal matters, such as the assessment of the validity of mining claims, mineral rights, and property agreements in the United States or elsewhere. Furthermore, the authors did not conduct any investigations of the environmental, social, or political issues associated with the DeLamar project, and are not experts with respect to these matters. The authors have therefore relied fully upon information and opinions provided by Integra and Mr. Edward Devenyns, Mineral Land Consultant for Integra, with regards to the following:

  • Section 4.2, which pertains to land tenure, including a Limited Due Diligence Review of the property prepared by Perkins Coie LLP (dated August 21, 2017) and a limited Updated Title Report review (dated March 8, 2022), as well as further information from Perkins Coie LLP dated March 2, 2018 and March 8, 2018; and

  • Section 4.3, which pertains to legal agreements and encumbrances.

The authors have relied fully upon information and opinions provided by Integra’s consultant, Mr. Richard DeLong of EM Strategies, Inc., an expert in environmental and permitting matters. Section 4.4, which pertains to environmental permits and liabilities, was provided by Mr. DeLong in communications via emails on September 25, 2017 (DeLong, 2017), July 17, 2019 (DeLong, 2019), and February 11, 2022 (DeLong, 2022a). Section 20, which discusses environmental permitting and related aspects of the project that was prepared by Mr. DeLong, was provided by Integra in a project communication via email dated February 25, 2022 (DeLong, 2022b).

The authors have fully relied on Integra to provide complete information concerning the pertinent legal status of Integra and its affiliates, as provided in Sections 1, 2, and 4, as well as current legal title, material terms of all agreements, and material environmental and permitting information that pertains to the DeLamar project, as summarized in Sections 1, 4 and 20.

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4.0PROPERTY DESCRIPTION AND LOCATION

The authors are not experts in land, legal, environmental, and permitting matters and express no opinion regarding these topics as they pertain to the DeLamar project. Subsections 4.2 and 4.3 were prepared under the supervision of Mr. Edward Devenyns, Mineral Land Consultant for Integra. Mr. Devenyns prepared a Limited Title Report on the unpatented claims dated August 15, 2017 and Perkins Coie LLP prepared a Limited Due Diligence Review dated August 21, 2017. On March 2 and March 8, 2018, Perkins Coie LLP provided MDA information concerning the Banner and Empire claims at Florida Mountain. A limited Updated Title Report review of the property was prepared by Perkins Coie LLP dated March 8, 2022. Mr. Richard DeLong of EM Strategies, Inc., an expert in environmental and permitting matters, prepared Section 4.4.

Integra owns 100% of the DeLamar project. All mineral titles are held or controlled by the DeLamar Mining Company (“DMC”), a wholly owned subsidiary of Integra.

The authors do not know of any significant factors or risks that may affect access, title, or the right or ability to perform work on the property, beyond what is described in this report.

4.1Location

Integra’s DeLamar gold-silver project is located in southwestern Idaho in Owyhee County, 80 kilometers (50 miles) southwest of the city of Boise, just west of the historical mining town of Silver City (Figure 4.1). The property is centered at approximately 43°00′48″N, 116°47′35″W, within the historical Carson mining district, and includes the formerly producing DeLamar silver-gold mine, which was last operated by the Kinross DeLamar Mining Company, a subsidiary of Kinross.

Figure 4.1 Location Map, DeLamar Gold – Silver Project

(modified from Hill and Lindgren, 1912; red numbers refer to 1912 mining districts)

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4.2Land Area

The DeLamar project includes 790 unpatented lode, placer, and mill site claims, and 16 tax parcels comprised of patented mining claims, as well as certain leasehold and easement interests located in Owyhee County, Idaho. In total, the property covers approximately 8,673 hectares (21,431 acres) owned or controlled by Integra (Figure 4.2) and occupies portions of:

  • Sections 30 and 31 of Township 4 South, Range 3 West;

  • Sections 28 through 36 of Township 4 South, Range 4 West;

  • Sections 25, 35 and 36 of Township 4 South Range 5 West;

  • Section 6 and 7 of Township 5 South, Range 3 West;

  • Sections 1 through 16 of Township 5 South, Range 4 West; and

  • Sections 1 through 3, 10, 11, 14, 15 and 22 of Township 5 South, Range 5 West, Boise Base and Meridian.

A listing of the patented and unpatented claims and leasehold interests that are included in the property is provided in Appendix A, Parts 1 through 6. Integra represents that the list of claims and leasehold interests in Appendix A is complete to the best of its knowledge as of the effective date of this report. Included in Appendix A, Part 1, are seven Idaho Department of Lands leases that have been issued to DMC.

DMC also owns mining claims and leases of State of Idaho lands located beyond the limits of the property described above. These landholdings are not part of the DeLamar project, although some of the claims are contiguous with those of the DeLamar and Florida Mountain claims and state leases.

Ownership of the unpatented mining claims is in the name of the holder (locator), subject to the paramount title of the United States of America, under the administration of the U.S. Bureau of Land Management (“BLM”). Under the Mining Law of 1872, which governs the location of unpatented mining claims on federal lands, the locator has the right to explore, develop, and mine minerals on unpatented mining claims without payments of production royalties to the U.S. government, subject to the surface management regulation of the BLM. Currently, annual claim-maintenance fees are the only federal payments related to unpatented mining claims, and these fees have been paid in full to September 1, 2022. The current annual holding costs for the DeLamar project unpatented mining claims are estimated at $138,680 (Table 4.1), including the county recording fees.

Other annual land holding costs, including county taxes for the patented claims and leased fee lands, and lease payments due to third-party claim owners, are listed in Table 4.1. The total annual land-holding costs are estimated to be $473,244.

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Figure 4.2 Property Map for the DeLamar Project

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Table 4.1 Summary of Estimated Land Holding Costs for the DeLamar Project

Annual Fee Type Amount
Unpatented Claims BLM Maintenance Fees $ 138,600
Unpatented Claims County Filing Fees $ 80
Estimated Holding Costs for Unpatented Mining Claims $ 138,680
Access, Pipeline, Land Agreement Fees $ 253,333
Owyhee County Patented Claims Taxes $ 5,849
Patented Claims Agreement Fees $ 48,100
State Lands Lease (annual rental and advanced minimum royalty payments) $ 27,282
Total Estimated Annual Holding Taxes and Fees $ 473,244

The reviews by Mr. Devenyns and Perkins Coie LLP have not identified any known significant defects in the title of the claims, and the authors are not aware of any significant land use or conflicting rights, or such other factors and risks that might substantially affect title or the right to explore and mine the property, based on the information provided by Integra and Perkins Coie LLP.

DMC holds the surface rights to the patented claims it owns and has leased, subject to various easements and other reservations and encumbrances. DMC has rights to use the surface of the unpatented mining claims for mining related purposes through September 1, 2022, and which it may maintain on a yearly basis beyond that by timely annual payment of claim maintenance fees and other filing requirements, and subject to the paramount title of the U.S. federal government. DMC holds surface rights to the areas it has under lease in accordance with the terms of each lease. These surface rights are considered sufficient for the exploration and mining activities proposed in this report, subject to regulation by the BLM and State of Idaho.

4.3Agreements and Encumbrances

On November 3, 2017, Integra announced that it acquired 100% of the DeLamar gold – silver project from a wholly owned subsidiary of Kinross for CAD$7.5 million in cash and the issuance of Integra shares. In addition, Table 4.2 summarizes further the agreements and encumbrances applicable to the property. Fees other than royalties associated with these agreements are included in the land-holding costs of Table 4.1.

In terms of royalties, 101 of the 284 unpatented claims acquired from Kinross are subject to a 2.0% net smelter returns royalty (“NSR”) payable to a predecessor owner (Table 4.2); this royalty is not applicable to the current project resources. There are also eight lease agreements that include 2% to 5.0% NSR obligations (referred to as Leases A through H in Figure 4.2, and Party A through G in Table 4.2) that apply to 33 of the patented claims and five unpatented claims. These claims are located within portions of Sections 1, 2, 4, 6, 11, and 12 of Township 5 South, Range 4 West; Sections 6 and 7 of Township 5 South, Range 3 West; Section 36 of Township 4 South, Range 4 West, and Section 31 of Township 4 South, Range 3 West, Boise Base and Meridian. Leases B and E apply to small portions of the DeLamar area (5% NSR to a maximum of $50,000) and Florida Mountain area (2.5% NSR to a maximum of $650,000) resources, respectively.

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The property also includes 1,561 hectares (3,857 acres) leased from the State of Idaho under seven separate State Mineral Leases that are subject to a 5.0% net smelter returns production royalty (Table 4.2), plus annual lease fees of $27,282 (Table 4.1) The state lease E600067 has an expiration date of February 28, 2028. The six other state leases have an expiration date of February 28, 2029. The State of Idaho leases include very small portions of both the DeLamar and Florida Mountain resources.

Kinross had retained a 2.5% NSR royalty that applies to those portions of the DeLamar area claims acquired from Kinross that are unencumbered by the royalties described above. The Kinross royalty applies to more than 90% of the DeLamar area resources; the royalty will be reduced to 1.0% upon Kinross receiving total royalty payments of CAD$10,000,000. The Kinross royalty has since been transferred to Maverix Metals Inc. (“Maverix”).

A total of approximately 20% of the current Florida Mountain resources and reserves are subject to one or more of the royalties described above. Figure 4.2 shows the areas subject to the royalties and lease agreements summarized in Table 4.2.

Table 4.2 Summary of Agreements and Encumbrances

(from Integra, 2021)

Owner Number of Claims or Lease Royalty
Maverix 183 unpatented claims and 13 tax parcels comprised of patented claims 2.5% NSR up to CAD$10M; then 1.0% NSR
Predecessor Owner 101 unpatented claims 2.0% NSR
State of Idaho 3,348 acres under seven separate Mining Leases 5.0% production royalty of gross receipts
Party A 1 patented claim 5.0% NSR to $50,000; then 2.5% NSR to a

maximum of $400,000
Party B 1 patented claim 5.0% NSR to a maximum of $50,000
Party C 2 patented claims 2.5% NSR
Party D 1 patented claim 2.5% NSR
Party E 9 patented claims and 1 unpatented claim 2.5% NSR to a maximum of $650K
Party F 12 patented claims 2% NSR to a maximum of $400K
Party G 7 patented claims 2% NSR
Party H 4 unpatented claims 2% NSR to a maximum of $80,000

Portions of the property are subject to a private land agreement, road access agreement, pipeline agreement, State of Idaho Easement Agreement and a BLM right-of-way agreement that include lands and certain rights within portions Sections 2, 3, 4, 7, 9, 10, 11, 14 and 18 of Township 5 South, Range 4 West, and Sections 11, 12, 13, 14, 23, 24, 25 and 26 of Township 5 South, Range 5 West.

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4.4Environmental Liabilities and Permitting

The 1977 – 1998 DeLamar open-pit mining operations included the DeLamar and Florida Mountain mining areas. The DeLamar area mine facilities, specifically the historical Sommercamp and North DeLamar open pits, incorporate essentially all the historical underground mining features (adits and dumps) in the vicinity. In the Florida Mountain area, many historical underground mining features remain to the north of the historical Florida Mountain open pits and waste rock dump, and several of these historical underground mining features are located within the DeLamar project, including collapsed adits, dumps, and collapsed structures. None of these features have water discharging to the environment.

The DeLamar mine has been in closure since 2003. Since 2003, the following reclamation and closure activities have been conducted on the DeLamar project:

  • Tailing pond de-watered and capped with clay and soil;

  • Four waste piles regraded and capped with clay and soil;

  • Heap-leach pad removed;

  • Much of the reclaimed surface includes an engineered cover consisting of two feet (61 centimeters) of compacted clay, 10 inches (25.4 centimeters) of non-acid generating run-of-mine (“ROM”) material, and 8 inches (20.3 centimeters) of suitable plant growth media;

  • The DeLamar mine facilities include three primary pit areas. These are the North DeLamar, Sommercamp – Regan (including North and South Wahl), and Glen Silver pits, which are partially backfilled and clay capped to allow for positive drainage;

  • The Florida Mountain mine facilities within the DeLamar project include the Jacobs Gulch waste-rock dump, which has been regraded and reclaimed, and the Tip-top, Stone Cabin, and Black Jack pits, which have been partly back-filled to allow for positive drainage;

  • The DeLamar mine is in the Closure Phase with the Idaho Department of Lands (“IDL”) and activities that focus on water management;

  • Water management includes collection of water at four primary collection and pumping stations referred to as Meadows, SP5, Spillway, and SP1. There are also two ancillary pumping stations at Adit 16 and SP14; and

  • The collection stations route water to a primary lime amendment facility and a smaller caustic-drip facility. Water passing through the lime amendment plant is routed to a storage pond and seasonally released at a nearby land application site (“LAS”).

The DeLamar project holds the following primary permits: two Plans of Operation (“PoO”), one with IDL and the BLM (PoO #248), and one with IDL (PoO #936). In addition, the DeLamar Mining Company holds a Cyanidation Permit from the Idaho Department of Environmental Quality (“IDEQ”), an Air Quality Permit from IDEQ, a Dam Safety Permit from the Idaho Department of Water Resources (“IDWR”), and a 2015 Multi-Sector General Permit (“MSGP”), Storm Water Permit, and a Ground Water Remediation Permit from the United States Environmental Protection Agency (“EPA”).

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Even though a substantial amount of reclamation and closure work has been completed at the site, there remain ongoing water-management activities and monitoring and reporting. The monitoring and reporting activities include: stream water quality and benthic, air quality, the LAS, and quality assurance and control. Water-management activities consist of year-round treatment with storage of treated water and discharge during the spring and summer months.

In January of 2017, Kinross submitted to IDL a reclamation bond reduction request, prepared by SRK Consulting (US) Inc. IDL responded in writing on April 24, 2017, indicating they had received the partial bond reduction request on March 29, 2017, and stated that they needed more time to complete the required site inspection prior to acting on the bond reduction request. On May 31, 2017, the IDWR issued a letter stating their relinquishment of any claims on the bond held by IDL. On June 19, 2017, IDL concurred with Kinross’ request for a $9,032,148 reduction in the bond. A reclamation bond of $2,778,929 remains with the Idaho Department of Lands (“IDL”) and a reclamation bond of $100,000 remains with the IDEQ. Additional reclamation bonds in the total amount of $589,144 have been placed with the BLM for exploration activities and groundwater well installation on public lands. There are also reclamation bonds with the IDL in the total amount of $86,900 for exploration activities on IDL leased lands.

As of the date of this report, Integra is conducting a drilling program on patented and unpatented mining claims in the DeLamar and Florida Mountain areas of the project. This drilling is being undertaken under a Notification from IDL, as well as two Notices filed with the BLM. The exploration program recommended in Section 26.0 includes proposed drilling in the Florida Mountain area of the project, as well as further drilling in the DeLamar area. This proposed work would necessitate a modification to the existing Notification for drilling in the DeLamar area, and a new Notification for Florida Mountain drilling performed on patented claims. A Notice would need to be filed with the BLM if any of the recommended drilling is undertaken on unpatented claims. Separate Notices would be filed with the BLM for each of the DeLamar and Florida Mountain areas of unpatented claims.

The authors are not aware of any significant factors and risks that may affect access, title, or the right or ability to perform work on the property, other than those discussed above.

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5.0ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE, AND PHYSIOGRAPHY

The information summarized in this section is derived from publicly available sources, as cited. Mr. Gustin and Mr. Weiss have reviewed this information and believe this summary is materially accurate.

5.1Access to Property

The principal access is from U.S. Highway 95 and the town of Jordan Valley, Oregon, proceeding east on Yturri Blvd. from Jordan Valley for 7.6 kilometers (4.7 miles) to the Trout Creek Road (Figure 5.1). It is then another 39.4 kilometers (24.5 miles) travelling east on the gravel Trout Creek Road to reach the DeLamar mine tailing facility and nearby site office building. Travel time by automobile via this route is approximately 35 minutes. Secondary access is from the town of Murphy, Idaho and State Highway 78 (Figure 4.1 and Figure 5.1), via the Old Stage Road and the Silver City Road. Travel time by this secondary route is estimated to be about 1.5 hours. Surface rights for access, exploration and mining are summarized in Section 4.2

Figure 5.1 Access Map for the DeLamar Project

(2022 property outline in green)

5.2Physiography

The property is situated in rolling to mountainous terrain of the Owyhee Mountains at elevations ranging from about 1,525 meters (5,000 feet) to 2,350 meters (7,710 feet) above sea level within portions of the De Lamar, Silver City, Flint, and Cinnabar Mountain U.S.G.S. 7.5-minute topographic quadrangles. Portions of the property are forested with second- or third-growth spruce, pine, aspen, and fir. Vegetation types include Douglas fir, juniper – mountain mahogany, sagebrush, mixed shrubs, and wyethia meadow communities.

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5.3Climate

The climate can be described as moderately arid in the lower elevations to mid-continental at the higher elevations, with warm summers and cold, snowy winters. MDA is unaware of published historical temperature and precipitation data for the Owyhee Mountains. Summer maximum temperatures can reach 32°C (90°F) and winter minimum temperatures can be as low as -40°C (-40°F) according to Integra site personnel. Precipitation at the mine site is believed to average about 50 centimeters (20 inches) per year, most of which occurs as winter snowfall. Snow cover at the upper elevations can be one to two meters (3 to 6 feet) deep. Mining operations have been demonstrated to be feasible year-round but do require snow removal equipment to maintain road access during the winter. Road access for exploration may be limited or interrupted by snow during December through April.

5.4Local Resources and Infrastructure

A highly trained mining and industrial workforce is available in Boise, Idaho, approximately 100 kilometers (62 miles) northeast of the project area. The project area is served by U.S. Interstate Highway 84 through Boise and by U.S. Highway 95 about 30 kilometers (18.6 miles) west of the site in southeastern Oregon. Mining and industrial equipment, fuel, maintenance, and engineering services and supplies are available in Boise, Idaho, as are telecommunications, a regional commercial airport, hospitals, and banking.

Housing, fuel, and schools are available in the nearby town of Jordan Valley, Oregon, which presently has a population of about 175 inhabitants. There are as many as a few dozen summer residents of the old historical mining town of Silver City, located about 8.5 kilometers (5.3 miles) east of the DeLamar mine, but few or no residents during the winter when road access is interrupted by accumulated snow.

An administrative office building with communications and an emergency medical clinic from the historical, late 20th century open-pit mining operation remain on site and in use. A truck shop and storage building also remain on site. The processing plant and facilities, crushing equipment, and assay laboratory have been removed from the property. Electrical power at the project site is delivered via a 69kV transmission line from the Idaho Power Company. Although the project area is generally hilly, flat areas are present and have served in the past for siting the processing plant and tailing storage areas. Developed water wells are present for mining and process requirements. Water required for the mining and process needs proposed in this PFS are discussed in Section 17.8 and Section 18.8. Areas for siting of the proposed waste-rock and tailings storage facilities, heap-leach pads, process plant, and energy needs are discussed in various portions of Section 18.

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6.0HISTORY

The information summarized in this section has been extracted and modified to a significant extent from Piper and Laney (1926), Asher (1968), Bonnichsen (1983), Thomason (1983), and unpublished company files, as well as other sources as cited. Mr. Gustin and Mr. Weiss have reviewed this information and believe this summary is materially accurate.

For clarity, this report will retain the term “De Lamar” to refer to the historical De Lamar underground mining operation of the late 19th and early 20th centuries and, consistent with official USGS topographic maps and place names, the historical De Lamar town site on Jordan Creek and De Lamar Mountain. According to Bonnichsen (1983), the present-day term “DeLamar” follows the usage of Earth Resources Company starting in the 1970s (see below). In this report, the term “DeLamar mine” refers to the open-pit mine and processing operation at De Lamar Mountain that began in the late 1970s.

6.1Carson Mining District Discovery and Early Mining: 1863 – 1942

Mining activity began in the DeLamar project area in May of 1863 when placer gold deposits were discovered in Jordan Creek, just upstream from what later became the town site of De Lamar (Wells, 1963 as cited in Asher, 1968). The placer deposits were traced up stream, beyond the DeLamar project area, and during the summer of 1863 the first silver-gold lodes were discovered in quartz veins at War Eagle Mountain, which is outside the portion of Integra’s property that is the subject of this report. This resulted in a rush of miners to the area and the initial settlement of Silver City. Several small mines at War Eagle Mountain were quickly developed with rich, near-surface ore. By 1866, there were 12 mills in operation (Piper and Laney, 1926). Grades decreased at depth and in 1875 the Bank of California failed, resulting in a loss of financial backing, which contributed to the closure of the mines by 1876. According to Lindgren (1900), cited in Bonnichsen (1983) and Piper and Laney (1926), an estimated $12 to $12.5 million was produced from the War Eagle Mountain veins from 1863 through 1875, or the equivalent of 600,000 to 625,000 ounces of gold. Silver-to-gold ratios of the ores during this period were on the order of 1:1 to 1:6 according to Piper and Laney (1926).

The general area of De Lamar, Florida Mountain, Silver City and War Eagle Mountain was known as the Carson mining district, which was larger than the current property controlled by Integra. There was only minor production from sporadic activity in the district at the War Eagle Mountain mines from 1876 through 1888, and some of the mines were never reopened. However, significant silver-gold veins were discovered during this time period at De Lamar Mountain and at Florida Mountain. Captain J.R. De Lamar founded the De Lamar Mining Company and was largely responsible for the development of important veins at the original, underground De Lamar mine, just to the south of Jordan Creek. De Lamar’s name was applied to the mine, the mountain, and the small mining town that was established on Jordan Creek.

In 1889, rich ore shoots were discovered in veins at the De Lamar mine area. De Lamar sold his interest to the London-based DeLamar Mining Company, Ltd. in 1901. Declining grades and increasing costs caused the closure of the De Lamar mines by 1914. An estimated total production value of precious metals of nearly $23 million was reported from the Carson district for the period 1889 – 1914 by Piper and Laney (1926). The De Lamar mine is believed to have produced approximately 400,000 ounces of gold and 5.9 million ounces of silver from a minimum of about 726,000 tonnes milled from 1891 through 1913, based on annual company reports (Gierzycki, 2004a). Mines in Florida Mountain are estimated to have produced a total of 133,000 ounces of gold and 15.4 million ounces of silver from 1883 to 1910 (Bonnichsen et al. undated, cited in Gierzycki, 2004a).

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Very little production took place in the Carson district until the 1930s, when gold and silver prices increased. Placer gold was recovered from Jordan Creek from 1934 to 1940, and in 1938 a 181 tonne-per-day flotation mill was constructed to process dumps from the De Lamar mine. The flotation mill reportedly operated until the end of 1942. In 1939, the Morrison-Knudson Company excavated a small open pit on the east side of Florida Mountain, but the operation was not profitable and was shut down in November of that year (Asher, 1968).

A summary of estimated annual production value for the entire district, including the DeLamar project, through 1942 is shown in Figure 6.1. Altogether, the district is believed to have produced about 1 million ounces of gold and 25 million ounces of silver from 1863 through 1942 (Piper and Laney, 1926; Bergendahl, 1964). Gierzycki (2004b) estimated a total district production of 0.6 million ounces of gold and 42 million ounces of silver for this period.

Figure 6.1 Estimated Annual Production Value, Silver City (Carson) Mining District 1863-1942

(from Asher, 1968)

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6.2Historical Exploration Since the 1960s

It is believed that mining properties in the De Lamar project area were largely inactive from 1942 until the mid-1960s. Anecdotal information suggests that the Sidney Mining Company and the Continental Materials Corporation (“Continental”) both engaged in diamond-core (“core”) drilling in 1966, but MDA has information only for the Continental drilling during this time. Continental’s holes were drilled to test veins down-dip from stopes of the old De Lamar mine (Porterfield, 1992).

During the late 1960s, the district began to undergo exploration for near-surface, bulk-mineable gold-silver deposits, but few records of the work are available. The Glen Silver Mining Company conducted core drilling in what later became either the Glen Silver or the Sommercamp area of the DeLamar project, but the exact locations of the drill holes are not known to MDA.

In 1969, the “Silver Group” was formed as a joint venture comprised of Earth Resources Company (“Earth Resources”), Superior Oil Company, and Canadian Superior Mining (U.S.) Ltd. The Silver Group acquired property in the De Lamar – Florida Mountain area and conducted geological mapping and sampling. Much of the early exploration work was carried out by Perry, Knox, Kaufman Inc. for Earth Resources, the operator of the project.

During 1969 and 1970, Earth Resources carried out trenching, sampling, and surface geological work, and drilled 39 conventional rotary drill holes at De Lamar Mountain. This resulted in the discovery of broad areas of near-surface silver-gold mineralization in the Sommercamp and Glen Silver zones, and what Earth Resources termed the North DeLamar zone. Following these discoveries, Earth Resources ramped up exploration and development drilling, and from about 1971 through 1976 at least 432 holes were drilled, mainly in the North DeLamar, Glen Silver, Sommercamp – Regan (including North and South Wahl), and Ohio areas (Figure 6.2). This drilling also included the first holes drilled at the nearby Sullivan Gulch and Milestone prospects, as well in the Florida Mountain area.

The Sidney Mining Company drilled eight core holes in the Sommercamp and North DeLamar zones in 1972. In 1974, Perry, Knox, Kaufman Inc. completed a feasibility study for the Silver Group with reserve estimates for an open-pit mining scenario at the Sommercamp and North DeLamar zones. In 1977, Earth Resources commenced operation of the DeLamar silver-gold mine with initial open-pit mining at the North DeLamar and Sommercamp zones (see Section 6.3 for a summary of the DeLamar mine production). In 1981, Earth Resources was acquired by the Mid Atlantic Petroleum Company (“MAPCO”), and Earth Resources continued to operate the DeLamar mine and exploration joint venture.

Earth Resources continued to explore the Sullivan Gulch, North DeLamar, Glen Silver and Florida Mountain zones between 1978 and mid-1984. Incomplete records show that at least 135 holes were drilled by Earth Resources in these areas of the property.

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Figure 6.2 Aerial View, Zones of Exploration and Mining Since 1969 within the DeLamar Area

(produced by MDA, 2019)

Note: North and South Wahl are included in what is referred to as the Sommercamp – Regan zone.

In September of 1984, the NERCO Minerals Company Inc. (“NERCO”) purchased MAPCO’s interest in the DeLamar project and became the operator of the joint venture. Less than a year later, in mid-1985, NERCO purchased the interests of the remaining joint venture partners and thereby attained 100% ownership of the project.

During 1985 through 1992, NERCO conducted extensive exploration and development drilling, as well as surface mapping and sampling. Drilling was focused mainly on expansion and definition of bulk-mineable mineralization at Florida Mountain, with significant amounts of drilling also completed at North DeLamar, Glen Silver, Sullivan Gulch, Town Road, and Milestone. Incomplete records indicate that a minimum of 1,594 holes were drilled by NERCO within the DeLamar project during this period.

NERCO was purchased by the Kennecott Copper Corporation (“Kennecott”), then a subsidiary of Rio Tinto – Zinc Corporation (“RTZ”), in 1993. Two months later in 1993, Kennecott sold its 100% interest in the DeLamar mine and property to Kinross.

Kinross continued exploration of the property while operating the DeLamar mine. A total of 338 exploration and development holes were drilled by Kinross in 1993 through 1997. Most of the drilling was focused on the Glen Silver, North DeLamar, and Florida Mountain areas of the project.

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In addition to the surface sampling, drilling, and geological work, several campaigns of geophysical studies were performed at various times in the project history.

Kinross ceased exploration work in 1997 and mining was halted at the end of 1998 due to unfavorable metal prices. In 1999, milling ceased and Kinross placed the DeLamar and Florida Mountain operations on care and maintenance. Mine closure activities commenced in 2003. Mine closure and reclamation were nearly completed by 2014, including removal of the mill and other mine buildings, and drainage and cover of the tailing facility.

The property continued to be in closure and monitoring from 2014 to 2017.

6.3Modern Historical Mining: 1977 through 1998

Total open-pit production from 1977 through 1998, including the Florida Mountain operation, is estimated at approximately 750,000 ounces of gold and 47.6 million ounces of silver (Gierzycki, 2004b). Although the mill reportedly continued to operate for some unknown amount of time in 1999, historical production records are only available to the end of 1998.

Earth Resources commenced open-pit operations and milling at the DeLamar mine in 1977. The mine initially operated five days per week with a target production of about 9,980 tonnes per day of ore and waste. Ore was processed by grinding in ball mills followed by agitated tank leaching with cyanide prior to precipitation with zinc dust. By the late 1980s, NERCO was mining ore and waste that totaled 21,772 tonnes per day and the mill processing capacity was 1,996 tonnes per day. At the time of the Kinross acquisition in 1993, the DeLamar mine was operating at a mining rate of 27,216 tonnes per day and a milling capacity of about 3,629 tonnes per day (Elkin, 1993). The DeLamar mine produced 421,300 ounces of gold and about 26 million ounces of silver from about 12.9 million tons mined from start-up in 1977 through to the end of 1992 (Table 6.1). Production during this period came from pits developed in the Glen Silver, Sommercamp – Regan, and North DeLamar areas.

Kinross commenced production at Florida Mountain in 1994, while continuing operations at the DeLamar mine, moving Florida Mountain ore to the DeLamar mill via an 8.4-kilometer (5.2 mile) haul road. Material was excavated from three open pits on the west side of the crest of Florida Mountain from 1994 through 1998. These were named the Stone Cabin, Tip Top, and Black Jack pits (Figure 6.3 and Figure 6.4). The Florida Mountain operation was formally referred to as the Stone Cabin mine in permitting and other documents. Gierzycki (2004b) estimated that 124,500 ounces of gold and 2.6 million ounces of silver were produced from the Stone Cabin mine in 1994 through the end of mining in 1998, based on an examination of files and company reports at the DeLamar mine

Mining in the Glen Silver – Sommercamp – North DeLamar areas continued simultaneously with the Florida Mountain operation. It has been reported that 625,500 ounces of gold and 45 million ounces of silver were produced from the Glen Silver – Sommercamp – North DeLamar areas over the entire life of mine from 1977 through 1998 (Gierzycki, 2004b).

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Table 6.1 DeLamar Mine Gold and Silver Production 1977 – 1992

(from Elkin, 1993)

Year Ore Mill Grade Bullion Poured
(short dry tons) Gold Silver total troy ounces
(oz/ton) (oz/ton) Gold Silver
1977 309,000 0.034 3.55 9,600 853,000
1978 637,000 0.031 3.78 18,100 1,872,000
1979 715,000 0.034 3.12 22,200 1,734,000
1980 780,000 0.031 2.53 22,100 1,534,000
1981 771,000 0.034 2.55 24,000 1,529,000
1982 738,000 0.036 2.77 24,300 1,589,000
1983 846,000 0.035 2.32 27,100 1,526,000
1984 784,000 0.023 2.83 15,500 1,742,000
1985 820,000 0.038 2.66 29,800 1,751,000
1986 849,000 0.035 2.52 27,700 1,713,000
1987 861,000 0.037 2.54 30,200 1,738,000
1988 830,000 0.033 2.34 32,000 1,738,000
1989 840,000 0.033 2.56 34,000 1,863,000
1990 829,000 0.037 2.04 30,400 1,374,000
1991 1,117,000 0.035 1.99 36,700 1,702,000
1992 1,156,000 0.035 2.01 37,600 1,820,000

Figure 6.3 Aerial View of the Florida Mountain (Stone Cabin Mine) Area

(produced by MDA, 2019)

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Figure 6.4 Photograph of the Reclaimed Florida Mountain (Stone Cabin) Mine Area

(view looking south-southeast)

6.4Historical Resource and Reserve Estimations

The estimates described in this subsection are presented herein as an item of historical interest with respect to historical open-pit mining and exploration at the DeLamar property. The historical estimations presented below are considered relevant because they represent an “ore reserve” that formed the basis of the initial open-pit mining, “reserves” estimated at the time of Kinross’ acquisition of the mining operations, and “resources” estimated at the time of closure of the open-pit mining operations. The classification terminology is presented as described in the original references, but these categories do not conform to the measured, indicated, and inferred mineral resource classifications as set out in NI 43-101 and the Canadian Institute of Mining, Metallurgy and Petroleum (the CIM Definition Standards). There is insufficient information for Mr. Gustin to understand how these historical categories differ from CIM Definition Standards. In addition, Mr. Gustin has not completed sufficient work to classify these historical estimates as current mineral resources or mineral reserves, and Integra is not treating these historical estimates as current mineral resources or mineral reserves. These historical estimates have been superseded by the current mineral resources described in this report and therefore they cannot be upgraded or verified as current mineral resources or reserves. Accordingly, these estimates are relevant only for historical context and should not be relied upon. The current mineral resources for the DeLamar project are discussed in Section 14.

The first reported historical “ore reserve” was presented in a 1974 feasibility study prepared by the Exploration Division of Earth Resources. A total of 4.124 million tonnes of “ore reserves” with average grades of 142.29 grams Ag/t and 1.58 grams Au/t, for about 18.8 million silver ounces and 210,000 gold ounces, were estimated for the Sommercamp and North DeLamar zones as shown in Table 6.2 (Earth Resources, 1974).

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At the time of the Kinross acquisition of the DeLamar operations and properties in 1993, the end-of-year 1992 reserves for the DeLamar mine area were estimated by Elkin (1993) at approximately 9.335 million tonnes with average silver and gold grades of 55.86 grams Ag/t and 0.72 grams Au/t, respectively (Table 6.2). Following the cessation of mining at the end of 1998 due to low metal prices, Kinross reported estimated resources and no reserves of 8.406 million tonnes with average silver and gold grades of 32.05 grams Ag/t and 1.25 grams Au/t, respectively (Table 6.2). The historical resources presented in Table 6.2 are based on the drill data available at the time of the estimations; the drill data are discussed in Sections 10.0, 11.0, 12.0, and 14.2.1.

Table 6.2 Historical Resource and Reserve Estimates

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Mr. Gustin has not done sufficient work to classify the historical estimates summarized in Table 6.2 as current mineral resources or mineral reserves, which are relevant only for historical context, and Integra is not treating these historical estimates as current mineral resources or mineral reserves. Mr. Gustin is unaware of the key assumptions, parameters, and methods used to prepare the historical estimates. Accordingly, these estimates should not be relied upon. The current mineral resources for the DeLamar project are discussed in Section 14.

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7.0GEOLOGIC SETTING AND MINERALIZATION

The information presented in this section of the report is derived from multiple sources, as cited. Mr. Gustin and Mr. Weiss have reviewed this information and believe this summary accurately represents the DeLamar project geology and mineralization as it is presently understood.

7.1Regional Geologic Setting

The DeLamar project is situated in the Owyhee Mountains, which are located near the east margin of the mid-Miocene Columbia River – Steens flood basalt province and the west margin of the Snake River Plain (Figure 7.1). The geology of various parts of the Owyhee Mountains has been described by Lindgren and Drake (1904), Piper and Laney (1926), Asher (1968), Bennett and Galbraith (1975), Panze (1975), Ekren et al. (1981), Ekren et al. (1982), and Bonnichsen and Godchaux (2006). As summarized by Bonnichsen (1983), Halsor et al. (1988), and Mason et al. (2015), the Owyhee Mountains comprise a major mid-Miocene eruptive center, generally composed of mid-Miocene basalt flows and younger, mid-Miocene rhyolite flows, domes and tuffs, developed on an eroded surface of Late Cretaceous granitic rocks. This Miocene magmatic and volcanic activity coincided with the regional Columbia River – Steens flood basalt event at about 16.7 to ~14.5 Ma (Mason et al., 2015).

Figure 7.1 Shade Relief Map with Regional Setting of the Owyhee Mountains

(from Mason et al., 2015)

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Note: OM = Owyhee Mountains; OP = Oregon Plateau; OIG = Oregon-Idaho graben; NNR = Northern Nevada Rift. Yellow shading shows the Columbia River – Steens flood basalt province; green shading indicates the Oregon Plateau underlain mainly by mid-Miocene silicic volcanic rocks. Red lines show eruptive loci and dike swarms; purple lines and ovoids are isochrons and silicic volcanic centers, respectively, with ages of silicic volcanism of the Oregon High Lava Plains and Snake River – Yellowstone provinces in Ma. Dark blue dashed and dotted lines are strontium isopleths. See Mason et al. (2015) for sources of data.

7.2Owyhee Mountains and District Geology

Five informal rock-stratigraphic sequences have been defined in the central Owyhee Mountains and the De Lamar – Silver City area (e.g., Ekren et al, 1981). From oldest to youngest these are the 1) Late Cretaceous Silver City granite; 2) mid-Miocene lower basalt; 3) mid-Miocene latite and quartz latite; 4) mid-Miocene Silver City rhyolite; and 5) mid-Miocene Swisher Mountain Tuff (formerly tuff of Swisher Mountain). The Silver City granite crops out near the crest and in the eastern part of the range (Figure 7.2), and it forms the pre-volcanic basement in the area. It has been described as mainly medium- to coarse-grained biotite-muscovite granodiorite to quartz monzonite and albite granite (e.g., Bonnichsen, 1983). It is considered to represent an outlying portion of the Idaho Batholith based on Late Cretaceous potassium-argon age dates, and similarities in composition, and mineralogy (Taubeneck, 1971; Panze, 1972). Integra’s district geologic map is shown in (Figure 7.2).

Figure 7.2 Geologic Map of the Central Owyhee Mountains

(from Integra, 2022; black lines are property outline)

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The Silver City granite is directly overlain by flows of the Miocene lower basalt, which have filled up to several hundreds of feet of relief on the granite. This demonstrates that the Silver City granite had been exhumed and underwent subaerial erosion by mid-Miocene time. The lower basalt is exposed in a northwest-trending band through the central part of the Owyhee Mountains (Figure 7.2) and consists of as much as 762 meters (2,500 feet) of flows of alkali-olivine to tholeiitic basalt that change upward to basaltic andesite and trachyandesite (Asher, 1968; Ekren et al., 1982; Bonnichsen, 1983; Thomason, 1983). As pointed out by Bonnichsen (1983), these basalts were erupted between 17 and 16 Ma, recalculated with modern decay constants from age dates of Panze (1975) and Armstrong (1975), and the lower part of the basalt sequence includes flows with distinctive large plagioclase phenocrysts, similar to flows of the Imnaha Basalt of the Columbia River Basalt Group.

Flows of latite and quartz latite overlie the lower basalt and in places directly overlie the Silver City granite (Thomason, 1983). The latite and quartz latite unit has a maximum thickness of about 549 meters (1,800 feet) (Panze, 1975).

The Silver City rhyolite (Asher, 1968) forms much of the central core of the Owyhee Mountains (Ekren et al., 1984) and consists of numerous individual and coalesced rhyolite flows and domes derived from local eruptive centers, as well as intercalated units of rhyolite ash-flow tuff (Panze, 1971; 1975; Thomason, 1983). Thomason (1983) estimated a composite thickness of as much as 457 meters (1,500 feet) for the sequence. Panze (1975) recognized a consistent succession of quartz latite, flow breccia and upper rhyolite that can be traced through the central Owyhee Mountains, and defined several vent areas and individual domes. More recent studies have shown that some of the individual quartz latite and rhyolite units consist of flow-layered, rheomorphic ash-flow tuffs of regional extent (Ekren et al., 1984).

The western and southern flanks of the Owyhee Mountains are capped by one or more cooling units of the Swisher Mountain Tuff, which overlies the Silver City rhyolite (Figure 7.2; Thomason, 1983; Ekren et al., 1984). To the west of DeLamar, the Swisher Mountain Tuff was emplaced at about 13.8 Ma as a regional sheet of unusually high-temperature rhyolite ash flows erupted from a vent area located near Juniper Mountain, about 64 kilometers south of De Lamar and Silver City (Ekren et al., 1984). Most of the unit is extremely densely welded and underwent post-compaction internal flowage (rheomorphic deformation), resulting in brecciated vitrophyres, contorted flow laminations and internal flow brecciation. In some places, however, eutaxitic textures and preserved pumice clasts provide evidence for the original ash-flow emplacement (Ekren et al., 1984).

Map patterns indicate the Owyhee Mountains have undergone incipient to minor amounts of mid-Miocene and younger regional extension. The principal faults recognized in the central Owyhee Mountains have normal displacements and primarily north-northwest orientations (Figure 7.2) approximately parallel to the Northern Nevada Rift (Figure 7.1). As stated by Bonnichsen (1983), “The attitude of the volcanic units generally ranges from subhorizontal to gently dipping, most commonly southwards. It is not clear if all the dips are due to initial deposition on uneven topography, or if some of the units have been rotated.

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7.3DeLamar Project Area Geology

7.3.1DeLamar Area

Earth Resources and NERCO geologists defined a local volcanic stratigraphic sequence in the DeLamar area based on geologic mapping and drilling. Mapping at various times benefited from exposures in the walls of the Glen Silver, Sommercamp – Regan, and North DeLamar pits. In addition to internal company reports, the geology of the DeLamar area has been documented in studies by Thomason (1983), Halsor (1983), Halsor et al. (1988), and Cupp (1989). These workers were involved with the exploration and operation of the project. The most concise and complete description of the local stratigraphic units and the mine area geologic setting was given by Halsor et al. (1988) and is presented here in Table 7.1. The Silver City granite is not exposed in the DeLamar area and has not been penetrated by drilling, although it is considered likely to underlie the Miocene rocks at depth.

Table 7.1 Summary of Volcanic Rock Units in the Vicinity of the DeLamar Mine

(modified from Halsor et al., 1988)

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The mine geologists considered the units above the lower basalt to be subunits of the Silver City rhyolite. However, the quartz latite (unit Tql, Table 7.1) has been correlated with the tuff of Flint Creek, a regional, high-temperature lava-like ash-flow tuff (Ekren et al., 1984).

Figure 7.3 shows the principal mineralized zones of the DeLamar project in relation to the DeLamar project outline, Figure 7.4 shows the surface geology of these mineralized zones, and Figure 7.5 shows a schematic geological cross section. Open-pits of the DeLamar mine were developed at the Glen Silver, Sommercamp – Regan, and North DeLamar zones. The Sullivan Gulch and Milestone zones have not been mined.

Figure 7.3 Land Position Map Showing Mineralized Zones

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Figure 7.4 Integra Generalized 2018 DeLamar Area Geology

(geology from Integra, 2019)

Note: Red outlines are schematic surface projection of the resource footprint; blue lines are faults. UTM grid NAD83, Zone 11; Y = North, X = East

Figure 7.5 Integra 2018 Schematic Cross-Section, DeLamar Area

(geology from Integra, 2019; line of section and rock unit legend shown in Figure 7.4)

Note: see Figure 7.4 for geology legend. UTM grid NAD83, Zone 11; X = East, Y = North, Z = elevation in meters.

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Mapping and drilling by Earth Resources and NERCO geologists has led to the interpretation that the mine area and mineralized zones are situated within an arcuate, nearly circular array of overlapping porphyritic and banded rhyolite flows and domes. These flows and domes overlie cogenetic, precursor pyroclastic deposits erupted as local tuff rings (Halsor, 1983; Halsor et al., 1988). Halsor (1983) interpreted the porphyritic and banded rhyolite flows and domes to have been emplaced along a system of ring fractures developed above a shallow magma chamber that supplied the erupted rhyolites, while Integra believes the rhyolites and latites were emplaced along northwest-trending structures as composite flow domes. The magma chamber was inferred to have been intruded within a northwest flexure of regional north-northwest trending Basin and Range faults (Figure 7.6).

Figure 7.6 Volcano-Tectonic Setting of the DeLamar Area

(showing land boundaries; modified from Halsor et al., 1988)

Core drilling in 2018 by Integra has facilitated the recognition of a unit of hydrothermally altered tuffaceous mudstone that is locally present between the porphyritic rhyolite and the overlying banded rhyolite as shown in Figure 7.5. This mudstone unit is up to 14 meters (46 feet) in thickness, strongly altered to clay, and includes fragmental volcanic layers of probable pyroclastic origin (Sillitoe, 2018; Hedenquist, 2018).

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7.3.2Florida Mountain Area

The geology of the Florida Mountain area has been described in general by Lindgren (1900) and Piper and Laney (1926). More detailed studies were carried out by Earth Resources and NERCO as documented by Lindberg (1985), Porterfield and Moss (1988), and summarized by Mosser (1992). The oldest stratigraphic unit is the Late Cretaceous Silver City granite, which is unconformably overlain by the mid-Miocene lower basalt to trachyandesite lavas. The granite and lower basalt are overlain by a sequence of andesitic volcanic-sedimentary and tuffaceous lacustrine rocks, which are in turn intruded and overlain successively by units of quartz latite, tuff breccia, and porphyritic rhyolite of the Silver City rhyolite (e.g., Lindberg, 1985). As at DeLamar, the tuff-breccia unit is interpreted as a near-vent pyroclastic unit erupted as a precursor to emplacement of the rhyolite flows and domes. Integra’s geologic map of the Florida Mountain area is shown in Figure 7.7.

In contrast to the DeLamar area, the Silver City granite crops out on the flanks of Florida Mountain and was extensively penetrated by workings of the historical underground mines. It was designated granite (Figure 7.7) by the Integra geologists. Field relations demonstrate the lower basalt flows partially buried an erosional, paleotopographic high of Silver City granite. Surface exposures and maps of the underground workings, as well as early drilling at Florida Mountain, led Lindberg (1985) to infer the granite forms a northeast-trending ridge beneath a relatively thin capping of quartz latite, tuff breccia, and one or more flows of rhyolite lava. Integra’s schematic cross section through Florida Mountain is shown in Figure 7.8.

The Earth Resources, NERCO and Integra geologists interpreted certain rocks at Florida Mountain to represent volcanic vents from which portions of the rhyolite flows and possibly tuffs were presumably erupted, and which later were important foci of hydrothermal activity, alteration, and mineralization (e.g., Porterfield and Moss, 1988; Mosser, 1992). However, exposures of rock units at Florida Mountain were generally poor prior to the start of mining by Kinross in 1994 as explained by Lindberg (1985), and the criteria used by the above authors to define the vent facies units and to delineate their geometries are not known to the authors. Moreover, most of the drilling at Florida Mountain was done by conventional rotary and RC methods, which can make outcrop-scale rock textural characteristics much more difficult, to impossible, to discern and correctly interpret.

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Figure 7.7 Geologic Map of Florida Mountain

(from Integra, 2022)

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Figure 7.8 Schematic Florida Mountain Cross Section (Looking Northeast)

(from Integra, 2022)

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7.4Mineralization

Numerous studies of the gold and silver mineralization in the DeLamar project – Silver City area have been conducted, beginning in the late 1860s. The most definitive studies and descriptions have been those of Lindgren (1900), Piper and Laney (1926), Thomason (1983), Halsor (1983), Halsor et al. (1988), and Mosser (1992). Mr. Gustin and Mr. Weiss have reviewed this information and believe it reasonably describes the mineralization as presently understood.

7.4.1District Mineralization

Precious-metal mineralization has been recognized in two types of deposits: within 1) relatively continuous, quartz-filled fissure veins, and 2) broader, bulk-mineable zones of closely-spaced quartz veinlets and quartz-cemented hydrothermal breccia veinlets that are individually continuous for only several 10s of centimeters laterally and vertically, and of mainly less than 1.3 centimeters in width.

Fissure Vein Mineralization

Mineralization mined from bedrock prior to 1942 was of the fissure vein deposit type. A concise summary of this type of mineralization in the Carson district was given by Bonnichsen (1983), as follows:

“Nearly all of the gold- and silver-bearing veins in the district strike north to northwest, following the main fault and dike trends, and are thought to be the same age….

Most of the veins are fissures filled with quartz, accompanied by variable amounts of adularia, sericite, or clay. A few have been described as silicified shear zones.”

At the De Lamar underground mine, the veins were as much as about 23 meters (75 feet) in width, but more commonly were 6 meters (20 feet) in width or less. Referring to veins in the Florida Mountain area, Bonnichsen (1983) went on to state:

“The veins are narrow, in most places only a few inches to a few feet wide, but persist laterally and vertically for as much as several thousand feet. Within an individual vein, the gold and silver ore occurs in definite shoots, generally with a moderate rake and somewhat irregular outline. The localization of ore shoots has commonly been attributed to the presence of cross-fractures, or, in one instance (Trade Dollar Mine), to the intersection of the vein with the granite-basalt contact. Some of the most productive veins in the district follow thin basaltic dikes.

All three major rock units, the Silver City granite, the lower basalt-latite unit, and the Silver City rhyolite, are cut by mineralized veins. Most of the production at War Eagle Mountain, Florida Mountain, and Flint was from veins in the granite, while at De Lamar all of the production was from the rhyolite.

Naumannite (Ag2Se) is the principal hypogene silver mineral and normally is accompanied by variable but subordinate amounts of aguilarite (Ag4SeS), argentite, and ruby silver as well as other silver-bearing sulfantimonides and sulfarsenides. Where interpreted to have been reorganized by supergene activity (Lindgren, 1900; Piper and Laney, 1926), the principal silver minerals are native silver, cerargyrite, and some secondary naumannite and acanthite. In both the hypogene and the oxidized and supergene-enriched portions of the veins, the principal gold-bearing minerals are native gold and electrum. Variable amounts of pyrite and marcasite, and minor chalcopyrite, sphalerite, and galena occur in some veins; the base metal-bearing minerals become more abundant at deeper levels.

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Quartz is the principal gangue mineral. Much is massive, but some has drusy or comb structure and a lamellar variety is locally abundant. This lamellar (or cellular or pseudomorphic) variety consists of thin plates of quartz set at various angles to one another (see photographs in Lindgren, 1900; Piper and Laney, 1926). Each plate consists of numerous tiny crystals that have grown from either side of a medial plane. Lamellar quartz has been interpreted as the replacement of preexisting calcite (or perhaps barite) crystals. Adularia commonly shows crystal outlines developed as open-space fillings.”

Calcite is reported to be present in only a few veins in the district, such as the Banner vein at Florida Mountain (Piper and Laney, 1926). Adularia is sparse in veins of the historical De Lamar mine, but is an abundant component of veins at Florida Mountain and War Eagle Mountain (Lindgren, 1900; Piper and Laney, 1926).

Potassium-argon age dates of volcanic units cut by veins, and dates on vein adularia concentrates, indicate that vein mineralization in the Silver City district was coeval with rhyolite volcanism at about 16 to 15 Ma (e.g., Panze, 1972; 1975; Halsor et al., 1988). More recent high-precision Ar40/Ar39 ages of adularia extracted from four samples of veins immediately outside of the project range from 15.42 ±0.07 Ma to 15.58 ±0.06 Ma (Aseto, 2012), in good agreement with the earlier studies.

Bulk-Mineable Mineralization

Zones of bulk-mineable mineralization have been recognized in the district only since the early 1970s. Mining of this type of mineralization has only occurred in the DeLamar project at both the DeLamar and Florida Mountain areas. Accordingly, this type of mineralization is described below in Section 7.5.1 and Section 7.5.2.

7.5DeLamar Project Mineralization

Current mineral resources discussed in this report are in the Florida Mountain area and the DeLamar area, which includes the Milestone prospect.

7.5.1DeLamar Area

The modern DeLamar open-pit mine area encompasses the historical De Lamar mine where fissure-vein mineralization was mined from 1889 through 1913. Mineralized shoots in two sets of fissure veins, the Main De Lamar and Sommercamp veins, were mined from what are now the Sommercamp – Regan and North DeLamar open-pit zones of Figure 7.4, as shown in Figure 7.9 at the 4th level (elevation 1,902 meters) (6,240 feet).

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Figure 7.9 Veins of the Historical De Lamar Mine, Elevation 6,240 Feet

(from Asher, 1968; based on Piper and Laney, 1926)

Note: the area of the above figure is entirely within the property boundary shown in Figure 7.3.

Bonnichsen’s (1983) summary of the DeLamar area vein mineralization is as follows:

“The main De Lamar section, at the site of the present-day North DeLamar pit…was 1,300 feet long in a northwest-southeast direction and up to about 300 feet wide, as measured on the No. 4 level (6,240 feet elevation). The section contained the Hamilton-Wilson No. 9 vein striking N. 25° W. and dipping 45°-66° W., and the 77 vein striking N. 62° W. and dipping 35° SW. These were connected by smaller veins and stringers. At lower levels the veins assumed steeper dips, 65 to 80 degrees being common. The 77 vein was the most important producer. The Sommercamp section, at the site of the present-day Sommercamp pit…was a zone about 300 feet across that contained ten interlinked veins striking N. 18° W. and dipping 65°-80° W.

These ore-bearing zones plunged 20 to 30 degrees southward. In both, the southern limit of the ore was a clay zone several feet thick with a shallow dip to the south. These clay zones were known as iron dikes to the miners and were interpreted to be the low-angle De Lamar and Sommercamp faults by Piper and Laney (1926), Asher (1968), and Panze (1975). However, the excellent exposure in the present-day open-pit mines has shown that these zones really are mainly the thick basal vitrophyric section of the banded rhyolite unit (Tbr) which has been hydrothermally altered. In the underground workings, much of the rich silver ore-the “silver talc”-was extracted where the veins abutted against the base of this clay zone. With its shallow dip, this zone formed the upper as well as the southern limit to mineralization in both sections of the mine.”

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An indication of the grades mined can be found in Piper and Laney (1926), where the 77 vein was reported to have been stoped from 1893 through 1908 with average grades mainly of 17.14 – 20.57 grams gold per tonne, and about 44.57 – 1,714 grams silver per tonne, over widths of 0.305 to 7.3 meters (1 to 24 feet). The overall width of the 77 vein was as much as about 23 meters (75 feet). During this period most of the production came from elevations above 1,786 meters (5,860 feet), but some stopes were as deep as the 12th level at 1,768 meters (5,800 feet). Although the 77 vein was found to persist to the 16th level at an elevation of 1,712 meters (5,617 feet), the lowest elevation of workings, grades were largely sub-economic below the 10th level and only a small amount of production came from the 12th level (Piper and Laney, 1926). As pointed out by Piper and Laney (1926), there was little underground exploration, and the development that was done did not consider the southerly plunge of mineralization.

In addition to the fissure veins, the bulk mineable type of mineralization has been delineated in four broad, lower-grade zones, two of which overlap and are centered on the Sommercamp and main De Lamar fissure veins. This type of mineralization has been described by Halsor et al. (1988) as follows:

“Low grade mineralization occurs in porphyritic rhyolite where closely spaced veinlets and fracture fillings provide bulk tonnage ore. Most of the veinlets are less than 5 mm in width and have short lengths that are laterally and vertically discontinuous….Locally, small veins can form pods or irregular zones up to 1 to 2 cm wide that persist for several centimeters before pinching down to more restricted widths. In highly silicified zones, porphyritic rhyolite is commonly permeated by anastomosing microveinlets typically less than 0.5 mm wide. Most of the minute veining displays well-defined contacts with the enclosing rock and in some instances veins can be seen to sharply cut phenocrysts. Still, in other zones, microveinlets are less distinct and difficult to distinguish from groundmass silicification.

Networks of high-density, quartz-free fractures are the sites for supergene mineralization. Major fractures generally trend north-northwest, but less prominent intervening and crosscutting fractures are present. Major fractures commonly have steep dips and show reversals in direction of dip vertically along faces. Fracture fillings commonly consist of thin coatings of goethite and jarosite but occasionally can be filled with seams of sericite and kaolinite up to several centimeters wide. Above the clay zone, veining is characterized by narrow, chalcedony-lined fractures of irregular extent.

In the Sommercamp pit, the principal ore zone in porphyritic rhyolite occurred beneath the clay zone as a distinct shoot striking north-northwest, dipping 40° E; and plunging 9½° SE. It was 27 m thick at the south end and thickened to 90 m at the north end. The ore-waste boundary at the base of the shoot was sharp with ore-grade material (>2 oz Ag) in the shoot abruptly dropping to waste across a single 1.5-m sample interval. The base of the ore shoot was remarkably planar but dipped 40° E as mentioned above. The top of the ore shoot was undulatory and more or less defined by the base of the clay zone over the porphyritic rhyolite. Generally, major mineralized shoots in the Glen Silver, North DeLamar, and Sullivan Gulch zones all plunge 10° to 15° to the southeast. Determining the plunge in the North DeLamar pit proved difficult due to a very complex cross faulting pattern.

Ore mineralogy is reported by Thomason (1983) and Barrett (1985). Naumannite (Ag2Se) is the dominant silver mineral and acanthite (Ag2S) and acanthite-aguilarite [(Ag2S)-(Ag4)(Se,S)2] solid solution are the second most abundant. Remaining ore minerals consist of lesser amounts of argentopyrite (AgFe2S3), Se-bearing pyrargyrite [Ag3Sb(S,Se)3], Se-bearing polybasite [(Ag,Cu)16Sb2(S,Se)11], cerargyrite [AgCI], Se-bearing stephanite [Ag5Sb(S,Se)4], native silver, and native gold and minor Se-bearing billingsleyite [Ag7(Sb,As)(S,Se)6], pyrostilpnite [Ag3Sb(S,Se)3] and Se-bearing pearceite [(Ag,Cu)16As2(S,Se)11]. Ore minerals are generally very fine grained; 65 percent of the minerals average 62μ in diameter, with the remainder averaging 200μ (Rodgers, 1980). Naumannite, the dominant silver mineral, commonly occurs as finely disseminated grains in quartz veinlets and within some fractures. It is also found as crystal aggregates growing on drusy quartz that lines vugs. Acanthite, the second most abundant silver mineral, occurs as anhedral blebs in quartz gangue and hydrothermal clays commonly associated with naumannite. It also is frequently present as a late-stage mineral coating drusy quartz in vugs…. Pyrite is the most widespread metallic mineral occurring in veins and altered country rock. Pyrite occurs along the edges of veins but also as coatings on some of the younger minerals. Polymorphic marcasite is commonly associated with pyrite, forming lath shaped crystals and anhedral aggregates surrounding pyrite. In some zones, marcasite is intimately intergrown in irregular clots with pyrite….

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Vein gangue minerals consist almost entirely of quartz, with minor amounts of mosaic intergrowths of adularia. Texturally, quartz can be divided into three varieties: (1) cloudy, massive, fine-grained quartz, (2) lamellar quartz, and (3) clear, crystalline, coarse-grained quartz…. Cloudy, fine grained quartz, including a chalcedonic variety, is the dominant type in veins and veinlets that constitute ore. This quartz is characterized by turbid anhedral grains (<0.005 mm) rich in solid inclusions.

The host rocks at DeLamar are pervasively altered. The tuff breccia is altered to an assemblage of quartz, illite, pyrite, and marcasite. The alteration of the principal host of mineralization, porphyritic rhyolite, is vertically zoned. The alteration assemblage is quartz, illite, pyrite, and marcasite and locally in the upper portions there are complex assemblages including jarosite, and mixtures of alunite, goethite, and kaolinite; hematite with kaolinite; and illite plus kaolinite (Thomason, 1983; Barrett, 1985). The latter style of alteration produces a very conspicuous glaring white rock that overlies the principal ore zones at DeLamar. The porphyritic rhyolite is overlain by a clay zone which consists of variable quantities of mixed layers of illite and montmorillonite clays with 5 to 7 vol percent euhedral pyrite in fine-grained aggregates or as crystals up to a few millimeters across. In less altered areas relic perlitic structure can be seen, demonstrating that the clay zone was a basal vitrophyre of the banded rhyolite. Above the clay zone, feldspar in the banded rhyolite is altered to kaolinite and the groundmass contains finely disseminated hematite, trace amounts of epidote, and patches of cryptocrystalline quartz. Sparse chemical data (Halsor, 1983) indicate that at least some of the DeLamar rocks were potassium metasomatized.

Scattered zones of breccia in the banded rhyolite occur most frequently near the base of the unit. These breccias crosscut flow layering, some ranging up to several meters in length by several decimeters in width. The breccias consist of close-packed angular fragments of flow-banded rhyolite in a chalcedonic matrix. The fragments show little rotation and this, together with the crosscutting nature of the breccias, suggests a hydrothermal origin and not primary features related to flow.”

The above description seems to have been based on the Sommercamp and North DeLamar mineralized zones. Mr. Gustin and Mr. Weiss have no information to suggest that the Glen Silver and the unmined Sullivan Gulch mineralization is different in a general sense. However, there is no indication that major fissure-vein mineralization was mined historically or encountered in exploration drilling in the Sullivan Gulch and Glen Silver zones, where relatively shallow drilling to date has intersected mineralization of the bulk mineable type.

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Based on Integra’s core drilling, the clay zone described above by Bonnichsen (1983) and Halsor et al. (1988) at least locally consists of the altered mudstone unit between the porphyritic and flow-banded rhyolites. The clay zone is interpreted as having acted as an important aquitard and barrier to upwelling hydrothermal fluids during mineralization (Sillitoe, 2018).

Samples from three drill-core intervals were studied with optical microscopy and x-ray powder diffraction methods at Hazen Research Inc. (“Hazen”) in 1971 (Perry, 1971). In addition to identifying some of the silver minerals recognized by Thomason (1983) and Halsor (1988), the Hazen study noted that gold occurs as native gold and in electrum. The gold grains were reported to be “blebs” that “rarely exceed 5 microns in size” intergrown with quartz, and within and on naumannite (Perry, 1971). Electrum was found as silvery, nearly white blebs less than 5 microns in size “locked in cerargyrite”.

The DeLamar area mineralization is situated stratigraphically below the Millsite rhyolite, which is reported to be little affected by hydrothermal alteration and is considered to be post-mineral in age (Thomason, 1983; Halsor et al., 1988).

7.5.1.1Milestone Prospect

A shallow, hot-spring setting has been described by Barrett (1985) for gold-silver mineralization at the Milestone prospect, about 1 kilometer northwest and along the strike of the Glen Silver zone (Figure 7.3). According to Gierzycki (2004b):

The ore lies at the base of a basalt-rhyolite contact in hydrothermal eruption-breccia with clasts of porphyritic rhyolite within a large zone of cherty silicification. It is capped at the surface by a sinter….Major ore minerals are naumannite, Se-rich pyrargyrite and gold.”

7.5.2Florida Mountain Area

Both fissure veins and the bulk-mineable type of mineralization are present at Florida Mountain and both have contributed to past gold and silver production. The veins cropped out intermittently near the crest and on the flanks of Florida Mountain, in some cases with lateral continuity of 1.6 kilometers (1 mile) or more, even though vein widths were usually only a few meters or less. Dips are reported to be 75° to 80° W, transitioning in their northern extents to steep east dips (Piper and Laney, 1926). A longitudinal section showing stopes of the Black Jack – Trade Dollar mine is presented in Figure 7.10.

The veins in Florida Mountain were mapped in greater detail in the 1970s and 1980s by Earth Resources, NERCO and later by Integra geologists (e.g., Figure 7.7), in part with the benefit of trenching and drilling. The most complete historical vein and geologic map that Mr. Gustin and Mr. Weiss are aware of is a NERCO map from 1989. The NERCO 1989 map shows a somewhat different, more detailed picture of the vein array than Piper and Laney’s 1926 map.

Mosser (1992) summarized the vein mineralization as follows:

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“…Mineralization is strongly controlled by NNW-trending faults, and to a lesser degree by arcuate and ENE structures. Host rocks display a definite influence on mineral distribution. Within the granodiorite and basalt, where most of the historic production occurred, the veins are narrow and tight. However, within the more reactive and permeable quartz-latite and rhyolite units, the mineralization is more disseminated so that significant bulk mineable potential exists…

The vein deposits are dominated by quartz and adularia gangue. Quartz occurs in a variety of forms in a definite paragenetic sequence….

Hypogene gold and silver mineralization varies little with depth across known levels and is dominated by electrum, acanthite, and the silver sulfo-selenide aguilarite….”

In the quartz latite and rhyolite, at least some of the veins branch upward into multiple narrow veins and vein-cemented breccia, separated by intensely altered rhyolite, to form sheeted vein and breccia zones as much as 6.1 meters (20 feet) or more in width. These broader sheeted vein and breccia zones comprise the bulk-mineable style of mineralization at Florida Mountain, particularly where adjacent fracture networks and flow bands in the rhyolite have been permeated with narrow, discontinuous quartz and breccia veinlets. Four such zones were described by Mosser (1992), referred to as the Tip Top, Stone Cabin, Main Trend (Black Jack), and Clark deposits. The mineralogy and paragenesis of the gold and silver mineralization are similar, if not the same, as that described for the fissure veins. Details of the mineralogy and a fluid inclusion study were presented by Mosser (1992). Information on the length, width, depth and continuity of mineralization is summarized in various parts of Section 14.

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Figure 7.10 Longitudinal Section of the Black Jack – Trade Dollar Mine

(from Piper and Laney, 1926)

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8.0DEPOSIT TYPE

Based upon the styles of alteration, the nature of the veins, the alteration and vein mineralogy, and the geologic setting, the gold and silver mineralization at the DeLamar project is best interpreted in the context of the volcanic-hosted, low-sulfidation type of epithermal model. This model has its origins in the De Lamar – Silver City district, where it was first developed by Lindgren (1900) based on his first-hand studies of the veins and altered wallrocks in the De Lamar and Florida Mountain mines. Various vein textures, mineralization, and alteration features, and the low contents of base metals in the district are typical of what are now known as low-sulfidation epithermal deposits world-wide. Figure 8.1, below, from Sillitoe and Hedenquist (2003), is a conceptual cross-section depicting a low-sulfidation epithermal system. The host-rock setting of mineralization at the DeLamar project is similar to the simple model shown in Figure 8.1, with the lower basalt sequence occupying the stratigraphic position of the volcano-sedimentary rocks shown below. The Milestone portion of the district appears to be situated within and near the surficial sinter terrace in this model.

Figure 8.1 Schematic Model of a Low-Sulfidation Epithermal Mineralizing System

(After Sillitoe and Hedenquist, 2003)

As documented by Lindgren (1900) and Piper and Laney (1926), many of the veins in the district contain distinctive boxwork and lamellar textures where quartz has replaced earlier crystals of calcite. These textures are now known to result from episodic boiling of the hydrothermal fluids from which the veins were deposited. Limited fluid inclusion studies of quartz from veins in the upper part of Florida Mountain by Mosser (1992) support the concept of fluid boiling and indicate fluid temperatures were in the range of 235°C to 275°C (455°F to 527°F). Salinities measured by freezing point depressions were apparently in the range of 0.25 to 2.1 equivalent weight percent NaCl, with a mean of about 0.8 equivalent weight percent NaCl (Mosser, 1992). Halsor et al. (1988) reported fluid temperatures from late-stage quartz in the DeLamar mine of about 170°C to 240°C (338°F to 464°F), with salinities of 2.8 to 3.8 equivalent weight percent NaCl. The temperature and salinity data, and evidence for fluid boiling are typical of the low-sulfidation epithermal class of precious-metal deposits world-wide.

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Many other deposits of this class occur within the Basin and Range province of Nevada, and elsewhere in the world. Some well-known low-sulfidation epithermal gold and silver properties with geological similarities to the DeLamar project include the past-producing Rawhide, Sleeper, Midas, and Hog Ranch mines in Nevada. The Midas district includes selenium-rich veins similar to, but much richer in calcite, than the veins known in the DeLamar project. At both DeLamar and Midas, epithermal mineralization took place coeval with rhyolite volcanism, and shortly after basaltic volcanism, during middle Miocene time.

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9.0 EXPLORATION

This section summarizes the exploration work carried out by Integra. Drilling by previous operators is summarized in Sections 10.2 and 10.3. Integra commenced drilling in 2018 on patented claims in the DeLamar area of the project and subsequently conducted drilling elsewhere at DeLamar as well as in the Florida Mountain area. Drilling conducted by Integra is described in Section 10.4 and was on-going as of the effective date of this report.

9.1Topographic and Geophysical Surveys

A Light Detection and Ranging (“LiDAR”) topographic survey of the DeLamar and Florida Mountain areas was completed late in 2017. Integra also commissioned SJ Geophysics Ltd., of Delta, British Columbia, to conduct an Induced Polarization and Resistivity (“IP/RES”) survey of six lines using the Volterra-2DIP distributed array system for a total of 22.4 line-kilometers (13.9 line-miles) in the DeLamar area late in 2017. The survey was extended with an additional 10 lines in 2018, bringing the total survey to approximately 40 line-kilometers (25 line-miles). The IP/RES lines were spaced at 300 meters (984 feet) and utilized a potential dipole spacing with intermediate current spacing of 100 meters (328 feet). The results are shown in Figure 9.1 and Figure 9.2.

Figure 9.1 Plan View of Resistivity from 2017 and 2018 IP/RES Surveys

(from Integra 2019, claim outline of 2019; 3D inversion elevation 1,600 meters)

Note: heavy black lines for “Surface Vein Projection” are schematic representations of historically mined mineralized structures; north is up.

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Figure 9.2 Plan View of Chargeability from 2017 and 2018 IP/RES Surveys

(from Integra, 2019; claim outline of 2019; 3D inversion elevation 1,600 meters)

Note: heavy black lines for “Surface Vein Projection” are schematic representations of historically mined mineralized structures; north is up.

9.1.12019 Airborne Magnetic Survey

A helicopter high-resolution magnetic survey of the DeLamar – Florida Mountain area was conducted in 2019 by New Sense Geophysics Ltd., of Markham, Ontario. The survey used 200-foot line separation at an average terrain clearance of 39 meters (128 feet). The term “high resolution” for this purpose implies a tight line spacing, stinger mounted magnetometer, sample rate of at least 50 Hz, low helicopter speed in dissected topography, and micro-leveling of the data. Basic processing of the 2019 data was done by the contractor and additional magnetic products including reduction to pole, various derivative products such as vertical derivative and analytic signal were prepared by Robert Ellis of Reno, Nevada, using Oasis Montaj software (www.seequent.com). Both conventional susceptibility inversion (Li and Oldenburg, 1996) and magnetic vector inversion (“MVI”) were used to generate a 3D voxel solid of the susceptibility distribution. The induced magnetic field direction at the time of the survey had an inclination of about 66.7° and a declination of about 13.6°. The direction of magnetization vector for high amplitude magnetic sources (i.e., Miocene basalts) defined in this inversion model varied from flat to +30° with declinations that between -100° and +100°. This confirms that remanent magnetization of the mafic rocks is present and the position and geometry of signatures with respect to the source locations in products like the reduction-to-pole and related products can be shifted. The amplitude component from the MVI inversion is the magnitude of the susceptibility accommodating remanence and induced magnetization and is referred to for convenience as susceptibility. This susceptibility model is used for the analysis of source locations of mafic and felsic intrusions, and intrusives into the granodiorite at Florida Mountain. The model was also used to better interpret structure.

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9.1.22020 Induced Polarization and Resistivity Surveys

Induced polarization (chargeability) and resistivity data were collected at DeLamar from nine east-west lines spaced 300 meters (984 feet) apart and totaling 29 line-kilometers (18 line- miles) in 2020. Zonge International of Tucson, Arizona collected the data. A dipole-dipole configuration was utilized. 2D inversion models of the data, including the 2018 and 2018 distributed array data, were done using TS2Dip (www.zonge.com/legacy/ModelIP.html). A 3D inversion of the data produced a marginally deeper solid with the sacrifice of lost resolution at shallow depths. Consequently, the 2D model sections were gridded to 3D voxel solids and used to extract sections shown with geology interpreted from surface mapping and drilling for DeLamar and Florida Mountain.

9.2Rock and Soil Geochemical Sampling

Integra conducted rock-chip and soil geochemical sampling at the DeLamar area in 2018. A total of 2,920 soil samples in the DeLamar area were collected at 50-meter (164-foot) intervals along lines spaced 300 meters (984 feet) apart, and 475 rock-chip samples were also collected.

During 2019 through 2021, Integra and contractor personnel collected 429 rock samples in the DeLamar, Milestone and Florida Mountain areas. Contractor personnel from Rangefront Geological (“Rangefront”) of Elko, Nevada collected 298 soil samples in the DeLamar/Milestone area in 2019. A total of 2,332 soil samples were collected from the Florida Mountain area by Rangefront in 2019.

9.3Geologic Mapping 2020 – 2021

Integra geologists carried out geologic mapping at a scale of 1:5,000 in 2020 and 2021. Approximately 6.25 square kilometers (2.4 square miles) were mapped in the DeLamar area. About 50 square kilometers (19.3 square miles) were mapped in the Florida Mountain area.

9.4Database Development and Checking

A major effort in updating the DeLamar and Florida Mountain drill-hole databases was undertaken by Integra. Geologists re-logged cuttings from almost 2,500 historical RC drill holes and the re-logging data was added to the databases. This program included logging of oxidation types as oxide, mixed and non-oxide, the data for which had never before been collected. Mr. Gustin used this logging to create detailed oxidation models for both resource areas. In addition, Integra extracted information on underground workings, groundwater and/or moisture level of samples, sample quality, and notes of down-hole contamination from approximately 2,200 historical paper geologic logs stored at the project site and entered this information into electronic spreadsheets. MDA then augmented the project resource databases using these spreadsheets.

Integra also completed an extensive comparison of the DeLamar and Florida Mountain drilling assays to the original paper laboratory assay records. First, all drill-hole intervals which were missing assay data were identified. The historical paper laboratory assay records were then searched for the corresponding drill-hole intervals. In most cases, the gaps in assayed intervals were found to be “No Sample” intervals, and the databases were updated with the No Sample designation if this was the case. If assays were found for these intervals, the data was added to the databases. Next, approximately every 10th sample interval in the databases was compared to the original paper records. This amounted to 7.5% of the Florida Mountain intervals and 9.7% of the DeLamar intervals. The drilled interval ‘from’ and ‘to’ depths, as well as the gold and silver assays for the interval, were compared to the paper records. The few discrepancies found were corrected with the entries recorded in the paper records and a field in each database was attributed with a record of the checking. This work was in addition to the verification work completed by the authors that is summarized in Section 12.0.

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9.5Cross-Sectional Geologic Model

Utilizing the updated databases, as well as available surface geology, Integra geologists constructed 100 hand-drawn cross-sections at 30-meter (98.4-foot) spacing through the DeLamar mine area. Cross-sectional lithology and structure were interpreted on each section using the down-hole data. Trends of mineralized zones were interpreted on each of the sections as well using the down-hole assays. While working on this modeling, conflicts in the geological coding of nearby holes were inevitably discovered. The resolution of these discrepancies often led Integra to update the lithologic codes in the project database with their own logging of the historical RC chips. Cross-sectional geological modeling was also completed at Florida Mountain prior to the updating of the databases described above. These cross sections and geologic information were summarized and interpreted as described in Section 14.2, 14.3, and 14.4. In essence, the significant results of Integra’s exploration and interpretations of the exploration information are summarized in Section 14.2, 14.3, and 14.4 where these results and interpretations are applied to the estimation of the current mineral resources in this report.

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10.0DRILLING

The drilling described in this section was performed in the DeLamar and Florida Mountain areas of the property. This drilling was completed by historical operators from the late 1960s through 1998, and by Integra commencing in 2018.

10.1Summary

MDA has records for a total of 337,268 meters (1,106,522 feet) drilled in 2,836 holes in the DeLamar and Florida Mountain portions of the property as summarized in Table 10.1. This includes Integra’s drilling through the end of 2020, i.e., the drill holes used in the current resource estimation. Drilling at the project has continued through 2021 to the effective date of this report, but these new drill holes are not considered herein.

Table 10.1 DeLamar Project Drilling Summary

Records of historical drilling are incomplete with respect to dates, drilling methods, drilling contractors, and types of drills used. As of the effective date of this report, MDA has documentation for 2,625 historical holes drilled in the DeLamar area, including the Milestone prospect, and the Florida Mountain area, for a total of 275,790 meters (904,823 feet). Table 10.2 summarizes the historical drilling by operator and year.

Of the historical holes for which the drilling method is known, 602 of the DeLamar area holes were drilled by RC, 438 by conventional rotary, and 60 were core holes. Seventy-four percent of the historical holes in the DeLamar area were vertical. At Florida Mountain, 961 of the historical holes were drilled by RC methods, 58 by conventional-rotary methods, and 46 by diamond core methods; less than 10% of the historical holes were vertical. None of the conventional rotary holes were angled in either area. A combined total of 106 holes were drilled using core methods for a total of 10,822 meters (35,505 feet), or 3.9% of the overall meterage drilled. The median down-hole depth of all historical holes in the DeLamar area is 91 meters (298.6 feet), and the median depth in the Florida Mountain area is 123 meters (403.5 feet). The aerial distribution of drill holes in the DeLamar area is shown in Figure 10.1. Historical drilling in the Florida Mountain area is shown in Figure 10.2.

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Table 10.2 Historical Drilling at the DeLamar and Florida Mountain Areas

10.2Historical Drilling – DeLamar Area

10.2.1Continental 1966

The earliest drilling that MDA is aware of was completed by Continental in the area of the 77 vein of the old De Lamar underground mine and now the site of the North DeLamar pit. A total of 1,378 meters (4,521 feet) were drilled in five inclined core holes, but MDA is unaware of what type of drill rig was used, core diameter(s), or the identity of the drilling contractor.

10.2.2Earth Resources 1969 – 1970

In 1969 and 1970, Earth Resources drilled 39 conventional rotary holes, for a total of 2,303 meters (7,555.9 feet, in the North DeLamar, Sommercamp, and Glen Silver areas. All of the holes were vertical. Harris Drilling was the contractor for most of the drilling, some of which was done with a Failing 1500 drill rig. Eklund Drilling of Elko, Nevada, drilled one of the holes using a Mayhew 2000 drill. MDA is unaware of the type(s) and size(s) of drill bits used.

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Figure 10.1 Map of DeLamar Area Drill Holes

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Figure 10.2 Map of Florida Mountain Area Drill Holes

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10.2.3Sidney Mining 1972

Sidney Mining drilled eight core holes in the Sommercamp and North DeLamar zones in 1972. MDA is unaware of the drilling contractor, type of rig, and core diameter(s) used for this drilling.

10.2.4Earth Resources ~1970 – 1983

Between as early as possibly 1970 and the end of 1983, Earth Resources drilled 465 holes. Five of these were core holes drilled in the DeLamar area in 1975 with Longyear 38 and Longyear 44 core drills operated by Longyear Drilling. Five core holes were also drilled in 1975 in the Glen Silver area by the same contractor using a Longyear 44 rig. The core diameter was HQ for all 10 core holes.

A total of 384 conventional rotary holes, for 659,701 meters (215,554.5 feet), were drilled during this period in the DeLamar, Glen Silver, Sommercamp – Regan, Town Road – Henrietta, Milestone, Ohio, Millsite, and Sullivan Gulch areas (Figure 6.2). All of these holes were vertical. Contractors at various times included: Justice Drilling using a Mayhew 1000 rig; and Eklund Drilling using G-15, Mayhew 1500, Mayhew 2000, and Gardner-Denver 1500 rigs. Harris Drilling used a Failing drill for 21 holes in 1973. Eklund also used an Ingersoll-Rand TH60 drill in 1979 and 1980, and apparently one of the holes drilled by this rig was a 183-meter (600-foot) vertical RC hole.

Earth Resources drilled an additional 70 vertical holes of unknown type during this period, for a total of 5,202 meters (17,067 feet).

10.2.5NERCO 1985 – 1992

Available records show NERCO drilled 691 holes during 1985 through 1992. These include 351 RC holes for a total of 37,093 meters (17,067 feet), seven conventional rotary holes drilled in 1986 for a total of 640 meters (2,099.7 feet), 36 core holes for 1,902 meters (6,240 feet), and 28,720 meters (94,225.7 feet) of drilling for which the drilling method is not known. 532 of the holes were drilled vertically.

During this period, drilling took place at various times at North DeLamar, Glen Silver, Sommercamp – Regan, Sullivan Gulch, Ohio, Town Road, the tailing area, and an area known as “Heap Leach”. The Sullivan Gulch holes were drilled in 1985 or later using RC methods. Twelve vertical RC holes were drilled at the Ohio area, but the rig type and contractor are not available. Six core holes were drilled in the Glen Silver area in 1986 with a Longyear 44 drill. After some point in 1987, all of NERCO’s drilling was done with RC methods. Tonto Drilling used an Ingersoll-Rand TH60 RC drill for some of the drilling in 1987 and 1989. An in-house Canterra RC drill was also used in 1989. Ponderosa Drilling was the contractor for 30 core holes drilled in the Heap Leach area in 1990, but the type of drill and core diameter is not known to MDA. The NERCO Cantera RC drill was also used for 19 holes drilled in the Ohio area in 1991, and 19 RC holes drilled in the Ohio and Town Road areas in 1992.

10.2.6Kinross 1993 – 1998

Kinross drilled 239 holes in the DeLamar area, and only six of these holes were drilled vertically. Kinross drilled 55 RC holes (4,491 meters) (14,734 feet) in 1993 in the North DeLamar, Glen Silver, and Sommercamp – Regan areas. The drilling contractor was Stratagrout and a Discovery drill was used. In 1994 and 1995, Kinross drilled 181 RC holes (16,624 meters) (54,540.7 feet) located in the North DeLamar, Glen Silver, Ohio, and Sommercamp – Regan areas. AK Drilling was the contractor for 19 of these holes, and Drilling Services was the contractor for at least six of the holes. Available records indicate only one 158-meter (518.4-foot) inclined RC hole was drilled in 1996, and two additional inclined RC holes, for a total of 91 meters (298.6 feet), are bracketed to have been drilled between 1995 to 1998.

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10.3Historical Drilling – Florida Mountain Area

10.3.1Earth Resources 1972 – 1976

During 1972, 1975 and 1976, Earth Resources drilled a total of 3,405 meters (11,171 feet) in 45 vertical rotary holes in the Florida Mountain area. This drilling was done by Eklund Drilling of Elko, Nevada, using 13.34-centimeter (5.25-inch) diameter hammer bits. A Gardner-Denver 15 rotary rig was used for the 1972 holes and a Mayhew 1500 drill was used for the 1975 – 1976 drilling. Samples were collected over 3.048-meter (10-foot) intervals, but MDA is unaware of any other specific drilling and sampling procedures and methods.

10.3.2ASARCO 1977

ASARCO drilled four vertical rotary holes in 1977 between DeLamar and Florida Mountain in an area that is not presently part of the land controlled by Integra. These holes total 579 meters (1,899.6 feet) drilled, and are part of the project database, but were not used to estimate the current mineral resources. Samples were assayed over 3.048-meter intervals (10-foot), but MDA is unaware of the drilling contractor, type of drill used, or the drilling and sampling procedures and methods.

10.3.3Earth Resources 1980

In 1980, Earth Resources drilled nine vertical rotary holes at Florida Mountain for a total of 651 meters (2,135.8 feet). Eklund Drilling and D. Allen Drilling were the contractors. A Midway and Ingersoll Rand TH100 drill were used, respectively, with 13.34-centimeter (5.25-inch) diameter hammer bits. Samples were collected over 1.524-meter (5-foot) intervals, but MDA is unaware of any other specific drilling and sampling procedures and methods.

10.3.4NERCO 1985 – 1990

NERCO drilled 898 exploration holes at and near Florida Mountain from 1986 through 1990, by far the largest amount of drilling by a single historical operator (Table 10.2). Thirty-six of the holes, for a total of 4,488 meters (14,724.4 feet), were inclined HQ-diameter (63 millimeter) (2.5-inch) core holes, with the remainder drilled by RC methods (11,729 meters) (38,481 feet). Twenty-eight of these RC holes were drilled vertically. Incomplete records show that 5 water wells, for a total of 475 meters (1,558.4 feet), were also drilled in 1988. At least one, and possibly all, of the water wells were drilled with a CP650 drill operated by “Allberry”. The authors are not aware of the drilling contractor or type of rig that was used to drill the core holes. In 1986, a total of 7,393 meters (24,255.3 feet) were drilled in 50 RC holes by Becker Drilling with a Drill Systems rig, but no other information is available on the specific methods and procedures used. MDA is not aware of the drilling contractors, rig types, and drilling methods and procedures used for NERCO’s RC drilling in 1987, 1988, 1989 and 1990.

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10.3.5Kinross 1995 – 1997

During 1995 through 1997, Kinross drilled a total of 9,901 meters (32,483.6 feet) in 99 RC holes in the Florida Mountain area. All but three of the 99 holes were inclined. Available records suggest that Drilling Services Company (“DSC”) of Chandler, Arizona, was the contractor for the three holes drilled in 1995, and that a TH100 drill was used. Dateline Drilling of Missoula, Montana, was the contractor for the 1996 drilling, which totaled 4,907 meters (16,099 feet) in 49 holes. In 1997, a total of 4,658 meters (15,282 feet) were drilled in 47 RC holes at Florida Mountain by AK Drilling of Ramsay, Montana, with a Foremost Prospector rig. For the Kinross drilling, samples were collected over 1.524-meter (5-foot) intervals, but MDA is unaware of any other specific drilling and sampling procedures and methods.

10.4Integra Drilling 2018 -2020

Integra’s drilling through 2020 is summarized in Table 10.3.

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Table 10.3 Integra Drilling Summary

10.4.1DeLamar Area Drilling 2018 – 2020

A total of 40,353 meters (132,391.7 feet) were drilled in 140 holes in various parts of the DeLamar area in 2018 through 2020. Approximately 43% of the holes and 55% of the meters were drilled with RC methods. The balance of the DeLamar area holes were drilled with diamond core, or with an initial RC “pre-collar” followed by a core tail. Only one of the 2018 and five of the 2020 DeLamar area holes was vertical, with the others inclined at angles of -45° to -85°.

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The RC drilling in 2018 and 2019 was conducted by Boart Longyear of Elko, Nevada using an MPD 1500 track-mounted drill. Bit diameters varied from 12.065 centimeters to 15.558 centimeters (4.75 to 6.125 inches). RC drilling was conducted wet; samples were passed through a rotating vane-type splitter to obtain samples generally in the range of 4.54 kilograms (10 pounds) to 9.07 kilograms (20 pounds) when dry. The RC samples were transported from the drill pads to the on-site logging and storage facility each day.

In 2018, the core holes were drilled by Major Drilling of Salt Lake City, Utah using LF90 track-mounted drill. HQ- and lesser PQ-size core was recovered with wireline methods that involved triple-tube coring.

The 2019 and 2020 core drilling at DeLamar was conducted by Boart Longyear of West Valley City, Utah using a track mounted LF90 core rig. PQ- and lesser HQ-size core was recovered with wireline methods and triple-tube coring.

The 2018 through 2020 drill core was placed in plastic core boxes by the drilling contractor and transported from the drill sites to Integra’s secure sample logging and storage area at the historical DeLamar mine site on a daily basis.

10.4.2Florida Mountain Area Drilling 2018 – 2020

In the Florida Mountain area, a total of 21,124 meters (69,304.5 feet) were drilled in 71 core holes (Table 10.3). These holes were inclined at angles of -45° to -75°. The drilling was performed by Major Drilling and Boart Longyear using LF90 track-mounted drills. HQ- and lesser PQ-size core was recovered with wireline methods that involved triple-tube coring. The drill core was placed in plastic core boxes by the drilling contractor and transported from the drill sites to Integra’s sample logging and storage area at the DeLamar mine.

10.5Drill-Hole Collar Surveys

Nearly all historical drill-hole collar locations were surveyed in local mine-grid coordinates by one or more dedicated mine surveyors. It is Mr. Gustin’s and Mr. Weiss’ understanding that the mine-grid coordinate system was established in the 1970s by Earth Resources’ surveyors. Mine-grid coordinate 100,000 East and 50,000 North is located at the surveyed Section corner between Sections 32 and 33 of Township 4 South, and Sections 4 and 5 of Township 5 South, on the hillside north of the De Lamar town site. The exact surveying procedures and type of equipment used to survey hole locations are not known to Mr. Gustin and Mr. Weiss. Surveyed hole coordinates were hand recorded in multiple copies of collar coordinate logbooks. The logbooks show that coordinates for 44 holes were “taken from maps“. These are from several different areas of drilling and are mainly the older holes drilled in those areas.

The x and y collar locations of Integra’s 2018, 2019 and 2020 drill holes were surveyed by Integra geologists using a Bad Elf GPS. The measured coordinates were then processed using the Natural Resources Canada website. Based on check surveys of post-processed Bad Elf GPS coordinates, Integra found the accuracy at the project to be less than one meter, usually considerably less. Elevations were assigned to each of the post-processed GPS x and y coordinates using the LiDAR data (see Section 9.1).

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10.6Down-Hole Surveys

None of the historical RC and conventional rotary holes in the DeLamar area are known to have been surveyed for down-hole deviations, while only 33 RC holes drilled in the Florida Mountain area have down-hole survey information in the database. Conventional rotary and RC drill holes can deviate significantly, in both dip and azimuth, with increasing deviations as depths increase, primarily in the case of inclined holes. It is therefore likely that deviations occurred in the historical drill holes at the DeLamar project, particularly at Florida Mountain, but, as is discussed below, this is not considered to be a material issue in the estimation of the current project resources.

Integra used a REFLEX EZ-GYRO EG0270 down-hole survey tool to measure down-hole deviation in the 2018, 2019 and 2020 drill holes. The instrument was operated by Integra personnel to survey the RC holes, and by Major Drilling and Boart Longyear drillers to survey the core holes. Azimuth, dip, and temperature were measured at 15.24-meter (50-foot) intervals. A few of the 2018-2020 drill holes were also surveyed down hole with an optical and acoustic tele-viewer system, although this information was not used in the project database.

10.7Sample Quality and Down-Hole Contamination

Down-hole contamination is always a concern with holes drilled by rotary (RC or conventional) methods. Contamination occurs when material originating from the walls of the drill hole above the bottom of the hole is incorporated with the sample being extracted at the bit face at the bottom of the hole. The potential for down-hole contamination increases substantially if significant water is present during drilling, whether the water is from in-the-ground sources or injected by the drillers. Conventional rotary holes, in which the sample is returned to the surface along the space between the drill rods and the walls of the drilled hole, are particularly susceptible to down-hole contamination, although these concerns are limited at the DeLamar project due to the shallow depths and vertical orientation of the rotary holes, and the fact that a significant quantity of the rotary data was mined out during the historical mining operations.

Some of the drill-hole logs reviewed by MDA were found to have notations as to the presence of water during drilling, as well as occasional comments concerning drilling difficulties and sample sizes. Integra therefore comprehensively compiled sample quality information from the historical drill logs, and this information, which includes logged notes on intersected groundwater and/or drill-injected fluids, was used by MDA in the modeling of project resources. For example, intervals for which down-hole contamination was noted or suspected by historical operators were evaluated in the context of surrounding holes, and when such intervals were deemed by MDA to have suspicious results, they were excluded from use in the resource estimation. Intervals noted as having poor recovery were also flagged and not used in the estimation of the project resources. Beyond the historical notations of possible contamination, MDA noted a few other historical drill intervals that likely experienced down-hole contamination, and these intervals were excluded as well.

Down-hole contamination is not a significant issue with the historical drilling at the DeLamar project due to the relatively shallow depths of these holes (median down-hole depths of 91 meters) (298.6 feet) for the mostly vertical holes in the DeLamar area and 123-meter (403.5-foot) median down-hole depths for the predominantly angled holes at Florida Mountain). Few historical drill holes at the DeLamar area intersected the water table, while none did at Florida Mountain. A few of the deeper Integra RC holes drilled at the DeLamar area, which penetrated to depths significantly below the water table, do have strong evidence of down-hole contamination, and these intervals were flagged and removed from use in estimation of the resources.

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10.8Summary Statement

There is a complete lack of down-hole deviation survey data for the historical holes in the DeLamar area database, and the Florida Mountain area database includes deviation data for 33 RC and four core holes. While the paucity of such data is not unusual for drilling done prior to the 1990s, the lack of deviation data contributes a level of uncertainty as to the exact locations of drill samples at depth. However, in the DeLamar area these uncertainties are mitigated to a significant extent by the vertical orientation of three-quarters of the drill holes, the generally shallow down-hole depths, and the likely open-pit nature of any potential future mining operation that is based in part on data derived from the historical holes. Such uncertainties, while still minor, are more pronounced in the Florida Mountain area, where about 80% of the historical holes were inclined, and the holes were generally slightly longer than those in the DeLamar area. In consideration of the fact that any potential future mining operation that would rely in part on the reliability of the historical drill data would entail open-pit methods, the potential inaccuracies in the locations of drill samples imparted by the lack of down-hole surveys is not considered to be a material issue.

Down-hole lengths of gold and silver intercepts derived from vertical holes, which were almost exclusively historical holes, can significantly exaggerate true mineralized thicknesses in cases where steeply dipping holes intersect steeply dipping mineralization, for example in portions of the Sommercamp area. This effect is entirely mitigated by the modeling techniques employed in the estimation of the current resources, however, which constrain all intercepts to lie within explicitly interpreted domains that appropriately respect the known and inferred geologic controls and mineralized thicknesses.

The overwhelming majority of sample intervals in the DeLamar and Florida Mountain databases have a down-hole length of 1.52 meters (5.0 feet). This sample length is considered appropriate for the near-surface style of mineralization that characterizes the current mineral resources at both the DeLamar and Florida Mountain areas.

Beyond the sample-quality issues discussed in Section 10.7, which were identified and the affected samples removed from use in the estimation of the project resources, Mr. Gustin is unaware of any sampling or sample-recovery factors that materially impact the accuracy and reliability of the drill-hole data, and he believes that the drill samples are of sufficient quality for the purposes used in this report.

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11.0SAMPLE PREPARATION, ANALYSIS, AND SECURITY

This section summarizes all information known to Mr. Gustin and Mr. Weiss relating to sample preparation, analysis, and security, as well as quality assurance/quality control procedures and results, that pertain to the DeLamar project. The information has either been compiled by Mr. Gustin and Mr. Weiss from historical records as cited, or provided by Ms. Richardson, a longtime employee at the mine. Ms. Richardson’s contributions to this section are derived from personal correspondences with Mr. Gustin and Mr. Weiss, an internal mine memorandum by Richardson (1985), and a recent informal summary document compiled at the request of MDA. In this section, conversions from metric to U.S. customary units of measure are limited to the first occurrence of that measurement.

11.1Historical Sample Preparation and Security

Mr. Gustin and Mr. Weiss are not aware of sample-preparation procedures or sample-security protocols employed prior to the start-up of open-pit mining operations in 1977, although further detailed reviews of historical documentation may yield such information in the future.

Elkin (1993) stated that sample preparation procedures at the mine laboratory had remained relatively constant up to the date of his ore-reserve report. Drill cuttings were split at the drill site to obtain samples weighing approximately 4.5 kilograms (10 pounds). When received at the mine laboratory, the samples were dried and crushed to -10 mesh. Splits of 150 milliliter (9.15 cubic inch) volumes were then pulverized to pulps with 90% passing 100 mesh. At the date of Elkin’s report, one-assay-ton (30-gram) (1.06-ounce) aliquots were taken from these pulps for assaying.

Mr. Gustin and Mr. Weiss are unaware of any specific sample-security protocols undertaken during the various historical drilling programs at the DeLamar project. However, approximately 75% of the drill data in the DeLamar area database and 98% of the holes in the Florida Mountain area are derived from drilling undertaken after the open-pit mining operations had initiated. It is very likely that the drilling and sampling completed during the mining operations was undertaken in areas of controlled access.

11.2Integra Sample Handling and Security

Integra’s RC and core samples were transported by the drilling contractor or Integra personnel from the drill sites to Integra’s logging and core cutting facility at the DeLamar mine on a daily basis. The RC samples were allowed to dry for a few days at the drill sites prior to delivery to the secured logging and core-cutting facility.

The 2018, 2019 and 2020 core sample intervals were sawn lengthwise mainly into halves after logging and photography by Integra geologists and technicians in the logging and sample storage area. In some cases, the core was sawed into quarters. Sample intervals of either ½ or ¼ core were placed in numbered sample bags and the remainder of the core was returned to the core box and stored in a secure area on site. Core sample bags were closed and placed in a secure holding area awaiting dispatch to the analytical laboratory.

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All of Integra’s rock, soil and drilling samples were prepared and analyzed at American Assay Laboratories (“AAL”) in Sparks, Nevada. AAL is an independent commercial laboratory accredited effective December 1, 2020 to the ISO/IEC Standard 17025:2017 for testing and calibration laboratories. The drilling samples were transported from the DeLamar mine logging and sample storage area to AAL by Integra’s third-party trucking contractor.

The soil samples were screened to -80 mesh for multi-element analysis at AAL. MDA has no other information on the methods and procedures used for the preparation of Integra’s soil and rock samples.

11.3Historical Sample Analysis – Prior to Commercial Open-Pit Mining Operations

Prior to the opening of the mine in April 1977, all gold and silver analyses of drill-hole samples consisted of fire assays completed by commercial laboratories, primarily Union Assay Office of Salt Lake City, Utah (“Union Assay”). This includes the core holes drilled by Continental in 1966 and Sidney Mining in 1972, as well as pre-mining Earth Resources drilling. Assay certificates from other commercial laboratories reviewed by Mr. Gustin and Mr. Weiss from this time period include those from Rocky Mountain Geochemical Corp. of Salt Lake City, Utah (“RMGC”) and Western Laboratories in Helena, Montana. Several holes were also found to have had samples analyzed by Earth Resources Naciamento Copper Mine Laboratory (“Earth Resources Lab”), which apparently was an internal laboratory in Cuba, New Mexico operated by Earth Resources. Mr. Gustin and Mr. Weiss know of no other details of the sample analyses performed prior to the beginning of mining operations in April 1977.

11.4Historical Sample Analysis – During Commercial Open-Pit Mining Operations

Upon initiating mining operations in April 1977, all ore-control (blast-hole) samples and most samples from exploration and development drilling were assayed at the DeLamar mine laboratory. Until approximately 1988, these in-house analyses were completed by MIBK atomic absorption (“AA”) methods (Porterfield and Moss, 1988). Gold was solubilized from 20 grams (0.705 ounces) of material using an unspecified method and then extracted from the solution using methyl isobutyl ketone (MIBK), with the gold concentration determined by AA. Approximately 60% of the historical drill holes in the DeLamar area database and 28% of those in the Florida Mountain area holes were drilled prior to 1988.

From approximately 1988 through to the end of the open-pit mining operations, all analyses by the mine laboratory were completed using standard fire-assay methods. Records reviewed by Mr. Gustin and Mr. Weiss reveal that some samples during this period were analyzed by Chemex Laboratories, Inc. of Reno, Nevada; RMGC; Union Assay; Legend Inc. of Reno, Nevada; Western Laboratories; and Earth Resources Lab. Union Assay and RMGC were most commonly used. According to Ms. Richardson, all gold and silver analyses were completed by fire assay with a gravimetric finish. The mine lab used silver inquarts to measure gold and silver gravimetrically.

Repeat fire assays by the mine laboratory of samples prior to 1988 that were originally analyzed by AA at the mine laboratory showed that the silver AA results were consistently lower than the fire assays, sometimes significantly lower; although fire-assay checks of the AA gold results were stated to have compared well. The mine laboratory staff believed that the understatement of the silver AA values was due to a relatively coarse grind in the sample preparation, which ultimately resulted in incomplete digestion of silver-bearing minerals prior to the AA analyses. Sometime in 1980, the mine instituted a much more systematic check-assay program, whereby sets of silver-mineralized samples from each mine area, as defined by mine AA analyses, as well as from certain ranges of mine benches within a mine area, were selected for checking by fire assay. The AA and fire-assay analyses were then compared by area, and a linear factor was determined that was used to mathematically increase the AA values for each area or set of benches analyzed. Factored silver values of blast-hole samples were used by the mining operation to determine waste from ore. Silver AA adjustment factors were also determined for each developmental drilling area until 1985, when it appears that factoring of the silver AA values ended.

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The systematic fire-assay check program was continuously monitored, with changes to the silver adjustment factors occurring frequently. Documents reviewed by Mr. Gustin and Mr. Weiss indicate that the factor was subject to modification as frequently as once monthly for each active mining or developmental drilling area. Ms. Richardson stated that the factoring of the blast-hole silver AA analyses worked well, as evidenced by the reported close agreement between mined grades determined by blast-hole data and head grades determined at the mill.

Because the Florida Mountain area was mined from 1994 to 1998, all gold and silver of blast holes, and most of the drill holes as well, were analyzed by fire assaying methods. According to Ms. Richardson, a silver inquart was added prior to fire assaying due to the generally low silver concentrations at Florida Mountain relative to the DeLamar area.

In 1997, Kinross also shipped 1,691 Florida Mountain RC drill intervals to Legend Inc. in Reno, Nevada, for sample preparation and assays of gold and silver. The samples were crushed to nominal 10 mesh, then split to obtain a 200-gram (7.05-ounce) sub-sample that was pulverized to nominal 200 mesh pulp. Gold and silver were determined on 30-gram (1.06-ounce) aliquots using fire-assay fusion with a gravimetric finish.

No further details of the sample analyses completed during open-pit mining operations are known to Mr. Gustin and Mr. Weiss.

11.5Integra Sample Analysis

The same principal analytical methods were used at AAL for both soil and surface-rock samples collected by Integra. Gold was determined by fire-assay fusion of 60-gram (2.12-ounce) aliquots with an inductively coupled plasma optical-emission spectrometry (“ICP”) finish. Silver and 44 major, minor and trace elements were determined by ICP and mass spectrometry (“ICP-MS”) following a 5-acid digestion of 0.5-gram (0.018-ounce) aliquots. Rock samples that assayed greater than 10 g Au/t were re-analyzed by fire-assay fusion of 30-gram (1.06-ounce) aliquots with a gravimetric finish. Samples with greater than 100 g Ag/t were also re-analyzed fire-assay fusion of 30-gram aliquots with a gravimetric finish. Some rock samples were analyzed for gold using a metallic-screen fire assay procedure.

RC samples from the 2018 and 2019 drilling were dried upon arrival at AAL’s Reno facility. The dry samples were crushed to a size of -6 mesh and then roll-crushed to -10 mesh. One-kilogram (2.205-pound) splits of the -10-mesh materials were pulverized to 95% passing -150 mesh. Sixty-gram aliquots of the one-kilogram pulps were analyzed at AAL for gold mainly by fire-assay fusion with an ICP finish. Silver and 44 major, minor, and trace elements were determined by ICP and ICP-MS following a 5-acid digestion of 0.5-gram aliquots. Samples that assayed greater than 10 g Au/t were re-analyzed by fire-assay fusion of 30-gram aliquots with a gravimetric finish. Samples with greater than 100 g Ag/t were also re-analyzed fire-assay fusion of 30-gram aliquots with a gravimetric finish. Selected RC samples were analyzed for gold using a metallic-screen fire assay procedure.

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Integra’s 2018, 2019 and 2020 core samples were prepared and assayed at AAL for gold, silver, and multi-elements using the identical methods used for Integra’s RC samples.

11.6Quality Assurance / Quality Control Programs

Quality Assurance / Quality Control (“QA/QC”) programs undertaken as part of the various exploration and development drilling programs of historical operators and Integra are described in this subsection.

11.6.1Historical Operators

Approximately 25% of the historical exploration and development holes in the DeLamar area and 4% of the holes in the Florida Mountain area were drilled prior to the initiation of open-pit mining and the use of the mine-site analytical laboratory. In this time prior to the mining operations, quality assurance/quality control (“QA/QC) procedures were employed to monitor Union Assay’s analytical results, but these QA/QC data, which exist in paper form have not yet been compiled by Integra. The analytical results of the mine laboratory were monitored by resubmitting samples to the mine laboratory for check assaying, but documentation of these check analyses is incomplete.

According to the 1974 historical feasibility study (Earth Resources Company, 1974), the Union Assay results obtained prior to the initiation of open-pit mining were checked by sending composites of Union Assay pulps, splits of drill core, and Union Assay coarse rejects to the following laboratories for sample preparation, where required, and assaying: Southwestern Assayers and Chemists in Tucson, Arizona; Skyline Laboratories in Denver, Colorado; Western Laboratories in Helena, Montana; Hazen Research in Golden, Colorado; and the Earth Resources Lab in Cuba, New Mexico. The various check samples were analyzed by either fire assay or atomic-absorption methods. An evaluation of this program summarized in the historical feasibility documents concluded that, “Some variation does exist between the different firms, and since all are generally quite reliable, it is really impossible to determine which one is the best; fortunately, the variations are within reason and appear to fall within a normal and acceptable range of difference.

The Elkin (1993) report indicates that repeat (check) assays were routinely run at the mine laboratory, which was confirmed by Ms. Richardson. Elkin reported that all samples with silver values in excess of 10 ounces per ton (343 g/t) or gold values greater than 0.1 opt (3.43 g/t) were resubmitted to the mine laboratory for check assay. Original sample pulps and splits from every fourteenth coarse sample were also resubmitted to the mine laboratory on a routine basis. Mr. Gustin and Mr. Weiss have not found detailed documentation of these check analyses, and therefore could not independently evaluate the results. Elkin also stated that duplicate samples were not being sent to outside laboratories at the time of his report.

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The mine lab also completed duplicate MIBK analyses and/or fire assays as a check on the lab’s original MIBK results. Samples with gold concentrations greater than 0.02 ounces per ton (0.7 g Au/t) and those within “geologically interesting zones” were fire assayed by outside commercial laboratories using 60-gram charges. The mine lab performed checks of the outside lab results, using fire assaying techniques on 30-gram charges. Porterfield and Moss (1988) reported that these checks verified the results of the commercial labs.

During 1997, Kinross shipped a total of 1,134 pulps of exploration RC drill samples from Florida Mountain to Legend Inc., in Reno, Nevada, for check assaying of gold and silver. The samples had apparently been crushed, split, and pulverized in the DeLamar mine laboratory. At Legend, the pulps were analyzed by fire-assay fusion with gravimetric finish using 30-gram aliquots. Further documentation, including the check-assay results, of this program have not been found.

11.6.2Integra

Coarse blank material, certified reference materials (“CRMs”), and RC field duplicates were inserted into the drill-sample streams as part of Integra’s quality assurance/ quality control procedures. The blank material consisted of coarse fragments of basalt that was inserted approximately every 10th sample. Commercial CRMs were inserted as pulps at a frequency of approximately every 10th sample.

CRMs. Integra purchased commercial CRMs (certified reference materials) for use in their drilling programs. The CRM pulps were inserted into the primary sample stream and analyzed with the drill samples. The results were used to evaluate the analytical accuracy and precision of the AAL analyses of Integra’s drill samples.

In the case of normally distributed data, 95% of the CRM analyses are expected to lie within the two standard-deviation limits of the certified value, while only 0.3% of the analyses are expected to lie outside of the three standard-deviation limits. Note, however, that most assay datasets from precious-metal deposits are positively skewed. Samples outside of the three standard-deviation limits are typically considered to be failures. As it is statistically unlikely that two consecutive analyses of CRMs would lie between the two and three standard-deviation limits, such samples are also considered to be failures unless further investigations suggest otherwise. All potential failures should trigger investigation, possible laboratory notification of potential problems, and possible reanalysis of all samples included with the failed standard result.

Table 11.1 lists the details of the CRMs that Integra has used at the DeLamar project for the drilling assays considered in this technical report.

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Table 11.1 Integra Certified Reference Materials

The AAL gold analyses of the CRMs inserted with the 2018-2019 drill samples met normal performance thresholds, with a moderate number of ‘failures’, although gold analyses of many either tended to have a low bias or clearly showed a low bias. Figure 11.1 shows a plot of the AAL gold analyses of CRM CDN-GS-P6A, which has a certified value of 0.738 g Au/t. While none of the analyses are ‘failures’, there is a clear low bias in the analyses in the time period of the central portion of the plot. This is typical of the AAL gold analyses of most of the CRMs in the 2018-2019 time period.

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Figure 11.1 CRM CDN-GS-P6A Gold Analyses – 2018-2019 Drill Programs

The AAL silver analyses of the eight CRMs used in the 2018-2019 drilling programs that have certified silver values in addition to certified gold values returned excellent results, with generally good precision and accuracy, leading to few failures and no bias, except for silver analyses of SN784, which show a high bias although without failures. Figure 11.2 shows typical results for AAL’s silver analyses of the CRMs.

Figure 11.2 CRM SN74 Silver Analyses – 2018-2019 Drill Programs

A total of 837 CRMs were inserted into the drill-sample streams submitted to AAL from the beginning of 2020 through late August 2021, including some holes drilled after the resource database cutoff. All of these were analyzed for gold, while 574 were analyzed for silver as well. The results are summarized in Figure 11.3 for gold and Figure 11.4 for silver, which show the results of all AAL analyses for all CRMs based on their standard deviations from the expected values.

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Figure 11.3 All CRM Gold Analyses – 2020-2021 Drill Programs

Figure 11.4 All CRM Silver Analyses – 2020-2021 Drill Programs

As shown in Figure 11.3 and Figure 11.4, the AAL analyses included 17 gold failures and two silver failures. Fourteen of the gold failures were to the high side, i.e., the AAL analyses exceeded the upper control limit of over three standard deviations, but nine of these failures were from a single CRM (CDN-GS-P5G). The AAL analyses of this CRM are biased high, and low-side failures were generated. Absent this problematic CRM, the failure rates for both gold and silver are less than 1%.

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Coarse Blanks. Coarse blanks are samples of barren material that are used to detect possible contamination in the laboratory, which is most common during sample preparation stages. In order for analyses of blanks to be meaningful, they must be sufficiently coarse to require the same crushing and pulverizing stages as the drill samples. It is also important for a significant number of the blanks to be placed in the sample stream within, or immediately following, a set of mineralized samples, which would be the source of most contamination issues. In practice, this is much easier to accomplish with core samples than RC.

Blank results that are greater than five times the lower detection limit of the relevant analyses are typically considered failures that require further investigation and possible re-assaying of associated drill samples. The detection limit of the AAL analyses was 0.003 g/t for gold and 0.020 g/t for silver, so blank samples assaying in excess of 0.015 g Au/t and 0.100 g Ag/t are considered to be potential failures that should be subject to review and possible action. Figure 11.5 shows a plot of the AAL analyses of the coarse blanks (y-axis) versus the gold values of the previous samples, which would be the likely source of any in-lab contamination.

Figure 11.5 Blank Gold Values vs. Gold Values of Previous Samples – 2018-2019 Drill Programs

Of the 915 AAL analyses of coarse blanks submitted with the 2018-2019 drilling programs, 14 exceeded the failure threshold, with blank assays ranging from 0.016 to 0.099 g Au/t. Only the highest value exceeds 0.050 g Au/t, and it is therefore the only potentially material failure.

There are 889 AAL silver analyses of the coarse blanks. Using the reported detection limit of 0.020 g Ag/t, 93% of the AAL analyses of the blanks are technical failures. However, the highest value of the blank analyses is 4.86 g Ag/t, which is not of a magnitude that would be material to the project. Less than 3% of the AAL blank analyses are greater than 1.0 g Ag/t. Possible explanations for the extreme failure rate include: (i) the coarse blank material was not barren with respect to silver; and (ii) the reported detection limit of the silver analyses is inaccurately low.

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As part of the 2020-2021 drilling programs, 849 coarse blanks were inserted into the drill-sample streams and each blank was analyzed for both gold and silver. The AAL analyses produced 11 gold failures and 266 silver results that exceeded the putative threshold of 0.1 g Ag/t; only seven of the silver analyses exceeded 1.0 g Ag/t, with a high of 6.58 g Ag/t. Figure 11.6 shows the coarse blank gold assays compared to the preceding sample gold assays.

Figure 11.6 Blank Gold Values vs. Gold Values of Previous Samples – 2020-2021 Drill Programs

RC Field Duplicates. RC field duplicates are secondary splits of original 1.52-meter (five-foot) samples collected at the RC rig simultaneously with the primary sample splits. Field duplicates are used to evaluate the total variability introduced by subsampling, including at the drill rig and in the laboratory (subsampling of the coarse rejects and pulps), as well as the variability in the analyses. Field duplicates should therefore be analyzed by the primary analytical laboratory.

Excluding pairs in which both the RC field duplicate and primary sample assays returned less-than-detection-limit results, there are a total of 1,708 pairs of gold analyses and 2,199 pairs of silver analyses, all from the 2018-2019 drilling programs (no RC drilling was undertaken in the two deposit areas in 2020-2021). Figure 11.7 is a relative-difference graph that compares the RC duplicate data to the primary samples.

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Figure 11.7 RC Field Duplicate Gold Results Relative to Primary Sample Assays

There is no bias in the data, suggesting that there were no material issues with the drill-site and all additional downstream sample splitting. The mean of the duplicates (0.239 g Au/t) is very close to that of the primary samples (0.242 g Au/t), and the mean of the relative differences (“RDs”) is 1%. The variability is well within an acceptable range, especially so for an epithermal deposit, with an AVRD of 14% that includes all data (no outliers were removed), and which decreases at higher grades (e.g., at a 0.2 g Au/t mean of pair cutoff, the average AVRD is 6%).

The silver field-duplicate data yield very similar results as for gold. There is no bias evident in the relative-difference graph, the means of the silver analyses of the duplicates and original samples are identical, the average RD is -1%, and the average AVRD is 19%, decreasing to 9% at a more relevant mean of pair cutoff of 15 g Ag/t. No outlier pairs were removed from these statistics.

11.7Summary Statement

None of the analytical laboratories used during historical exploration and mining operations mentioned in this section were certified, as the formal certification process used today had not yet been implemented. Mr. Gustin and Mr. Weiss are not familiar with Western Laboratories or the Earth Resources Company internal laboratory, and the laboratories of Hazen Research and Southwestern Assayers and Chemists were not commonly used for routine assaying by the mining industry. However, historical documents reviewed by Mr. Gustin and Mr. Weiss indicate that Union Assay and, to a lesser extent, RMGC were the primary commercial laboratories used by all operators prior to Kinross, and these were independent commercial laboratories that were widely recognized and used by the mining industry at that time.

Documentation of the methods and procedures used for historical sample preparation, analyses, and sample security, as well as for quality assurance/quality control procedures and results, is incomplete and in many cases not available. It is important to note, however, that the historical sample data were used to develop and operate a successful commercial mining operation that produced more than 400,000 ounces of gold and 26 million ounces of silver. Mr. Gustin and Mr. Weiss are therefore satisfied that the historical analytical data are adequate to support the current resources, interpretations, conclusions, and recommendations summarized in this report.

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Integra’s sample preparation and analyses were performed at a well-known certified laboratory, and the sample security and assurance/quality control procedures are judged to be adequate.

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12.0DATA VERIFICATION

12.1Drill-Hole Data Verification

The drill-hole databases that support the estimations of the DeLamar and Florida resources are comprised of historical data and subsequent information derived from Integra’s 2018 through 2020 exploration programs. The historical portions of the databases were created by Mr. Gustin using the historical DeLamar mine digital database files for each of the two resource areas of the project.

The DeLamar area resource database includes information derived from 1,550 historical holes and 127 Integra holes. As the first step in verifying the historical data, a total of 235 of the historical holes were randomly chosen and their hole-collar coordinates, hole orientations, and gold and silver analytical information were compared to extensive, original, paper documentation in the possession of Integra. The database for the Florida Mountain area includes information from 1,107 historical drill holes, of which 169 were similarly checked, and 109 Integra holes that were all subjected to some level of checking. The results of this work, as well as other forms of data verification, are discussed below.

12.1.1Collar and Down-Hole Survey Data

DeLamar Area. Drill-hole collar location information was found in the historical documentation for 157 of the 235 holes selected for auditing. The locations of two holes were found to have substantially different locations in the project database compared to the paper records; it remains unclear as to which of the two sources is more accurate. A third hole had an 18-meter (59-foot) difference in elevation with the paper records, but the database elevation matches the project topography and was therefore deemed to be more accurate. All other location discrepancies were due to the rounding of surveyed locations documented in paper records to the nearest foot (0.31 meters), or the truncation of surveyed decimals in the mine-site database. These discrepancies, which Mr. Gustin considers immaterial to the resource estimation, may reflect the perceived accuracy of the original drill-collar location data.

There were no down-hole deviation data in the original mine-site database files. Ms. Richardson stated that no down-hole surveys were completed on conventional rotary or RC holes, which predominate the historical holes drilled at DeLamar. Six of the audited holes were core holes, but no deviation data were found in the paper records for these holes. Azimuth and dip records of the hole collars do exist, however, and no discrepancies were found between the historical paper records and the database.

The collar and hole-deviation surveys of approximately 25% of the Integra holes drilled at the DeLamar area were audited by comparing the information provided to MDA by Integra with original electronic files of the deviation surveys; no discrepancies were found.

Florida Mountain Area. Original x-y-z collar location data were found for 74 of the holes chosen for auditing. Three of these were found to have significant x-y discrepancies due to an updated survey location found in the historical records that was not entered into the original mine-site database. However, the holes as located in both the historical mine database and the updated survey information lie to the south of the current mineral resources. No discrepancies were found in the azimuth and dips of the audited holes.

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In addition to the auditing of the database values, several holes at both the DeLamar and Florida Mountain areas were found whereby the hole collars were significantly above pre-mining topography and/or the assay results (and often logged lithologies) were not consistent with those of nearby holes, suggesting the hole locations were not properly represented in the historical databases. In the few cases where historical documentation could not resolve these issues, the holes were flagged in the databases and not used in subsequent resource estimations.

The drill-hole collar locations and down-hole surveys of approximately 25% of the holes drilled by Integra at Florida Mountain were checked in a similar manner as described above for DeLamar. No discrepancies were found.

12.1.2Assay Data

Historical Assays. Historical paper records, including copies of original assay certificates, handwritten mine-lab assay sheets, and, to a lesser extent, handwritten assay values included on geologic logs, were used to audit the database gold and silver assay values from the historical holes. Documentation was found for 154 of the 235 historical holes selected to be audited in the DeLamar area database, and this led to the checking of 9% of all sampled and assayed historical intervals in the database. Discrepancies between the MDA database and paper records that are unrelated to the treatment of lower-than-detection-limit results or unanalyzed intervals were found in only nine of the 7,758 sample intervals audited, and less than half of these discrepancies were considered material. As part of this verification process, analytical data from a total of 195 historical sample intervals were found that were not included in the original database, and these data were added to the resource database.

Historical back-up data for the gold and silver values of 141 of the holes selected for auditing from the Florida Mountain area were found, representing 13% of the historical holes in the database and 12% of the historical sample intervals. A sequence error was found in which gold and silver values for one sample interval were repeated in the next sample interval, and the following eight gold and silver values were shifted down one sample interval (1.52 meters) (5.0 feet). The affected intervals are very low grade, except for a single 0.41 g Au/t value. In addition to this sequence error, one apparent transcription error was found, whereby the mine-site database had a value of 1.81 oz Ag/ton (62 g Ag/t) versus a value of 0.813 oz Ag/ton (30 g Ag/t) on the original assay sheet. These discrepancies were corrected.

Analytical data for 41 historical sample intervals in two holes drilled at Florida Mountain were found and added to the current project database as a result of the auditing.

During the auditing of the historical databases, certain gold analyses were found to be lacking in precision. These fire assays, primarily by independent commercial labs used by some of the earliest operators at the project, reported gold values in increments of 0.005 oz/ton (0.17 g/t). This lack of precision is particularly problematic at grade ranges of potential mining cutoffs, and especially so for low-cost open-pit operations such as is envisioned for the DeLamar project. A total of 11,197 DeLamar area gold analyses and 888 Florida Mountain area analyses were identified as being low precision; these sample intervals were flagged in the databases and were not used in the estimation of the project resources.

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The following discussion summarizes statistical analyses of various duplicate datasets Mr. Gustin compiled from the historical mine-site databases following the verification and related corrective work discussed above.

The mine databases have up to three mine-lab analyses for certain sample intervals, although the nature of the material re-assayed (e.g., pulps, coarse rejects, field duplicates) is not known. All available duplicate mine-lab analyses were compiled by Mr. Gustin and the duplicate datasets were evaluated.

The data for 1,762 drill samples for which both primary silver fire assays and second silver fire assays were performed by the mine lab, and both analyses were not below the detection limit, were examined. These data are summarized in Figure 12.1, which is a relative-difference graph. The graph shows the percentage difference (plotted on the y-axis) of each duplicate assay relative to its paired primary-sample analysis by the mine lab. The relative difference (“RD”) is calculated as follows:

Positive RD values indicate that the duplicate-sample analysis is greater than the primary-sample assay, while a negative value indicates the duplicate analysis is lower. The x-axis of the graph plots the means of the silver values of the paired data (the mean of the pairs, or “MOP”) in a sequential, but non-linear, fashion. The red line shows the moving average of the RDs of the pairs, which provides a visual guide to trends in the data that can aid in the identification of potential bias. A total of 108 pairs characterized by unrepresentatively high RDs have been excluded from the graph. In this and subsequent graphs, metal grades are shown in ounces-per-ton, honoring the units of the original analyses.

Figure 12.1 Repeat Mine Lab Silver Assays Relative to Original Mine Lab Assays

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Figure 12.1 suggests a high bias of low magnitude in the duplicate silver results relative to the original assays over most of the grade range of the data. The mean of duplicate analyses is 0.613 oz Ag/ton (21.0 g Ag/t), which is 4% higher than that of the original results (0.588 oz Ag/ton; 20.2 g Ag/t), and the average RD of the pairs is +2% (the average RD can be an approximate measure of the degree of bias, although one must be wary of the statistical effects of pairs with anomalously high RDs). The mean of the absolute value of the RDs (“AVRD”), a measure of the average variability exhibited by the paired data, is quite high at 73%, suggesting that the duplicate analyses were not completed on original-sample pulps. At a MOP cutoff of 1.0 oz Ag/ton (34.3 g Ag/t), the mean of the duplicate analyses of the 196 pairs is 5% higher than the original analyses, the average RD is +6%, and the mean AVRD drops to 16%. It should be noted that the high bias in the duplicates relative to the original analyses is present in what is an overall a relatively low-grade silver dataset, and the magnitude of the high bias over the majority of the grade range is low.

A similar dataset for 1,837 pairs of gold fire assays, after removal of 15 pairs that exhibit extreme variability, yields identical means (0.013 oz Au/ton; 0.45 g Au/t) for the duplicate and original analyses, an average RD of +1%, and a mean AVRD of 26%. The grades in this dataset are much more representative of the mineralization of interest than the silver duplicate data presented above.

Various check analyses of the original mine-lab assays were performed by various commercial, or “outside”, laboratories, primarily Union Assay and RMGC. Excluding 25 outlier pairs and all pairs in which the original and check assays were less than the detection limits, a total of 696 pairs of silver fire assays were evaluated. The nature of the material sent to the outside labs for analysis (pulps, coarse rejects, or field duplicates) is not known, nor is the identity of outside lab that performed the check analyses known, although it is believed that Union Assay completed most of them. These unknowns hinder the analysis. However, the mean of the outside lab duplicates (0.676 oz Ag/ton; 23.2 g Ag/t) is 7% lower than the mean of the original mine lab analysis for the complete dataset, and 8% lower at a cutoff of 1.0 oz Ag/ton (34.4 g Ag/t). The relative difference graph of the data (Figure 12.2) indicates that this discrepancy is largely caused by the prevalence of high-variable pairs having low values for the outside lab relative to the mine lab. Once again, it is important to note the relatively low-grade nature of the dataset. The moving-average line is of limited use in this case due to the effects of the numerous high-variability pairs.

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Figure 12.2 Outside Lab Silver Assays Relative to Original Mine Lab Assays

Only 28 outside lab fire assays for gold were found that were also assayed by the mine lab. The mean of the outside lab analyses for this limited dataset is 0.005 oz Au/ton (0.17 g Au/t), while the mine lab assays averaged 0.006 oz Au/ton (0.21 g Au/t).

As discussed in Section 11.0, the historical exploration and development drill-hole samples were variably analyzed for gold and silver by fire assay and AA methods, and for a period of time the mine-lab silver AA values were factored to account for incomplete sample digestions. The historical DeLamar and Florida Mountain databases that supported the open-pit mining operations documented the various types of analyses, with multiple analytical types commonly completed on a single sample interval. The databases also included “FFAU” and “FFAG” fields that were comprised of the gold and silver values, respectively, used for all mine-site purposes including reconciliations and historical estimations of resources and reserves. The FFAU and FFAG values prioritized fire assays completed by the mine site or outside laboratories over mine-lab AA analyses. The factored AA silver values were included in the FFAG field, while the original, unfactored AA silver analyses were also retained in the mine-site databases.

Mine lab AA silver analyses were reported to have been systematically low. Figure 12.3 compares data from 4,378 pairs of mine-lab fire assays and mine-lab AA analyses. A clear systematic bias is evident, whereby the AA analyses are lower than the paired fire assays, which is consistent with the mine staff’s observations. The mine site attributed this to incomplete digestions of silver minerals in the AA analyses. In an attempt to account for the digestion problem, the mine lab used the fire-assay data to factor the AA results for use in the mining operations. While the results of the relative difference graph were expected, this was not necessarily the case for the relatively constant magnitude of the low bias. This constancy of the low bias is seen visually in the relative-difference graph, and it is evidenced statistically. The mean of the AA analyses is 22% lower than the fire-assay mean for all data, as well as at several MOP cutoffs inspected. The average RD also is more-or-less constant at approximately -30% for all cutoffs examined. No original or factored AA silver analyses were used in the estimation of the project resources.

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Figure 12.3 Mine Lab Silver AA Analyses Relative to Mine Lab Silver Fire Assays

Figure 12.4 compares mine-lab gold AA analyses with mine-lab fire assays of samples from the same intervals.

Figure 12.4 Mine Lab Gold AA Analyses Relative to Mine Lab Gold Fire Assays

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All 4,797 pairs are shown, including many pairs with very high variability (the average AVRD is 323%). While the mean of the AA analyses for the entire dataset is 17% lower than that of the fire assays, the means are identical for all MOP less than 0.1 oz Au/ton (3.43 g Au/t). The mean of the AA analyses for the 111 pairs with MOP > 0.1 oz Au/ton is 45% lower than the mean of the fire assays. This demonstrates that the difference in the means for the entire dataset is due solely to differences in the highest-grade portion of the data. Accordingly, higher-grade AA gold values may be understating the actual grades of the samples. The AA gold values in sample intervals for which no fire assay gold analyses are available, which represent 29% of the historical gold assays in the DeLamar resource database and 24% in the Florida Mountain database, were accepted for use in the estimations of the current resources. All factored mine-lab AA silver values were removed from the resource databases. Unfactored mine-lab AA silver analyses that remain in the databases due to the lack of fire assays for those sample intervals were flagged and not used in the resource estimations, as these analyses demonstrably understate silver grades. In contrast, MDA’s analysis of mine-lab AA gold analyses found that they agree well with paired fire assay data up to a grade of 0.1 oz Au/ton (0.343 g Au/t), but at higher grades the AA gold analyses tend to be lower than the paired fire assays. While fire assays were prioritized over the AA gold analyses in the resource databases, the AA analyses are used for sample intervals lacking fire assays.

Integra Assays. Integra provided MDA with a complete assay compilation for all holes drilled in 2018 through 2020. The sample numbers in these files were then linked by MDA to original laboratory digital assay certificates to comprehensively validate the Integra assay tables by comparing all Integra assays to the original laboratory certificates. No discrepancies were found during this checking other than in a few cases where MDA chose to use a certain analytical method when multiple methods were available, and the method chosen by MDA differed from that in the Integra compilation.

12.1.3Integra Data Verification

In addition to MDA checking of historical data using historical records, Integra independently verified the accuracy of the ‘from’, ‘to’, and gold and silver assay values of every 10th sample interval, using the MDA’s audited database (see Section 9.4). The very few discrepancies found by Integra were then corrected in the resource databases, and this work also led to further checking of the surrounding sample intervals.

At MDA’s request, Integra also compiled data relevant to sample quality from historical drill logs and logged oxidation state using historical chipboards stored at the mine site.

12.2Additional Data Verification

In addition to the more structured verification procedures discussed above, extensive verification of the project data, with an emphasis on the historical data, was undertaken throughout the process of the resource modeling. The careful work involved in the explicit modeling of the gold and silver mineralization within the context of the project geology led to ad-hoc checking of the accuracy of a variety of data, such as hole locations, hole orientations, drill-hole lithologic attributes, and specific gold and/or silver assays. For example, during Integra’s cross-sectional geologic modeling, and Mr. Gustin’s modeling of the mineralization, historical holes were identified as having lithologic and assay information that was strongly at odds with adjacent holes. While paper survey records supported the database locations in some cases, a judgment was made that the holes’ locations must be inaccurate, and these holes were therefore excluded from use in the resource estimation. Many individual historical assays, as well as assays within entire mineralized intervals, were questioned and then confirmed by paper records and, in some cases, corrected in the project database as a result of working closely with the data during modeling.

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The Integra drilling provided another important component to the verification of the historical data. Integra’s ongoing drilling programs led to repeated updates of the resource databases. After each batch of new Integra drill data was added, the data were compared to the existing gold, silver, and lithologic modeling as it was being updated to reflect the new data. Where the Integra drill data penetrated areas at both the DeLamar and Florida Mountain areas that were previously modeled on the basis of historical data, the addition of the Integra data did not lead to material changes to the volume or grade of the gold and silver mineralization. This detailed work with the Integra drill data in the context of the historical information played a critical role in the validation of the historical data.

To further verify the historical data in a more quantitative manner, the 2021 DeLamar and Florida Mountain resource models were compared to the 2019 models prior to updating the new models with additional density data (which led to slightly higher densities as compared to the 2019 models). In order to further assure the comparisons were meaningful, the optimized pits used to constrain the 2019 resources were used to tabulate the 2021 models. The resulting tabulation of the DeLamar 2021 model, using the 2019 resource pits and cutoff parameters, yielded 2% more tonnes and essentially identical gold and silver grades as compared to the 2019 model. At Florida Mountain, the 2021 model has essentially identical tonnes at a gold grade that is 0.01 g/t lower than 2019 and a very similar silver grade. The DeLamar 2021 resource model incorporated data from an additional 50 holes (primarily core) drilled by Integra as compared to the 2019 modeling, while the 2021 Florida Mountain modeling was updated by incorporating the results from an additional 44 Integra core holes. The very close correspondence of the 2021 and 2019 resource models within the identical volumes of the 2019 resources pits demonstrates that the Integra data are consistent with the historical data, which supports the visual validations of the historical data discussed above.

12.3Site Inspection

Mr. Weiss visited the project site for three days, on August 1 – 3, 2017, accompanied and assisted by Ms. Kim Richardson of Jordan Valley, Oregon. Ms. Richardson is a geologist who joined the DeLamar mine staff in 1980 and eventually held the positions of Senior Mine Geologist, Mine Superintendent, and Mine General Manager before leaving the operation in 1997. Mr. Weiss reviewed the property geology, exposures of mineralized rocks within and near the still accessible open pits, and areas of historical exploration drilling peripheral to the open pits, in both the DeLamar and Florida Mountain areas. Historical exploration data on file at the DeLamar mine-site office was reviewed, including geologic maps and cross sections from various areas, mainly dating to the late 1980s.

Mr. Weiss attempted to verify historical drill-hole collar locations peripheral to the open pits. Nearly all historical drill sites external to the pits and waste dumps have undergone reclamation since closure activities began in 2003. Seven drill collars were found in the Sullivan Gulch and Ohio areas. Metal tags marked with the hole numbers were found at a few of the collars, but none of these were legible. Nevertheless, the eight collar locations were recorded with a hand-held Garmin GPS-62 receiver in UTM WGS84 projection in case that the holes can be identified in the future.

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Mr. Gustin visited the project site on October 16 through October 18, 2018, October 15, 2020, and October 27, 2020. All principal areas of mineralization at the DeLamar and Florida Mountain areas were visited in the field during these visits, as well as exploration areas both on the project (Town Road – Henrietta) and north of the project on other Integra-controlled lands. Numerous altered and mineralized areas throughout the project and adjacent areas were visited, open-pit walls were examined, and mineralized intervals from multiple core holes were closely inspected. An RC rig was also visited, and all project procedures related to the RC and core drilling programs, data collection, and data storage were reviewed. Where appropriate, recommendations were provided to the Integra technical team.

Mr. Dyer completed a site visit on October 27,2020. Mr. Dyer verified existing facilities and general topography. This is much of the extent of the data used by Mr. Dyer for mine planning and economic evaluation. Mr. Dyer concludes that the data used are suitable for this report.

Mr. Welsh performed an initial site appraisal for Integra in late June of 2019. He was escorted around the site by Messrs. Tom Jordan and Tim Arnold of Integra. The principal sites visited had been previously identified as potential heap-leach pad sites between DeLamar and Florida Mountain. As Mr. Welsh travelled between Jordan Valley and the mine site, he viewed a potential tailings location approximately 2.5 miles west of the mine in an undeveloped meadow. Generally, the access around the property was good via pickup truck and several legacy mining and exploration roads were observed that were available for future site evaluation and characterization. The sites visited included the legacy mine pits at Florida Mountain, the reclaimed waste-rock storage facilities in Rich Gulch, the north facing ridges between Florida Mountain and DeLamar mine area, and the accessible areas around the operating water treatment plant.

In August 2020, Mr. Welsh toured potential tailings and heap-leach facility sites with Mr. Tim Gerkin, a geologist. On the afternoon of August 28, 2020 Mr. Welsh visited the Slaughterhouse Gulch drainage between Jordan Creek and the headwater area. This trip resulted in changing the recommended locations of the heap-leach facility and the tailing storage facility from the PEA due to geotechnical and perceived permitting constraints under then new Idaho Interim Regulations for ore processing with cyanide.

Mr. Welsh also attended a pre-bid on-site meeting for prospective consultants on October 27, 2020. A brief trip to an overlook was made to view the proposed locations of the heap-leach pads and process area.

Mr. Nopola also completed a site visit on September 24, 2020 and reviewed general site and existing pit slope conditions.

12.4Independent Verification of Mineralization

No samples were collected from the DeLamar project for verification purposes by the authors. Gold and silver production from the historical underground mines and more recent open-pit operations is well documented in both private historical records and publicly available documents. In the opinion of Mr. Gustin and Mr. Weiss, independent sampling for the purposes of verifying the DeLamar and Florida Mountain mineralization is not needed.

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12.5Metallurgical Data Verification

Mr. McPartland visited the DeLamar project site on January 17, 2019. Most of the metallurgical test data used for the PFS was from testing conducted under the direct supervision of Mr. McPartland at McClelland. Where subcontractors were used to generate metallurgical or mineralogical test data, Mr. McPartland reviewed the reports submitted by the subcontractors concerning that testing. Mr. McPartland also reviewed reports concerning historical metallurgical testing, where available, though that historical testing was used only as general background information for the project metallurgy. Mr. McPartland considers the metallurgical test data to be of sufficient quality to be used in this report.

12.6Data Verification for Mine Engineering and Geotechnical

Mr. Welsh reviewed geotechnical sampling and laboratory testing of hollow stem auger samples collected under his supervision for characterization of foundation conditions for the Slaughterhouse Gulch tailing storage facility and the heap-leach pad area in 2021. He also reviewed previous work performed at the property for NERCO as reported in LaFronz et al. (1989), but used data collected in 2021 for his analyses. Mr. Welsh concludes that the 2021 data are suitable for use in this report.

Mr. Nopola reviewed all geotechnical data collected under his supervision in 2020-2021 including laboratory rock core sampling and laboratory strength test results, discontinuity orientations obtained from photogrammetric surveys of existing slopes, and geotechnical core logging of exploration core. This data was used the characterize the materials expected to be encountered in the Florida Mountain and DeLamar open pit mine slopes. The geotechnical study also considered historical data described by NERCO (1987) and Golder Associates (1989). Mr. Nopola and/or engineers under his direct supervision thoroughly reviewed these historical reports and the historical data was both validated and supplemented with data collected in 2020/2021. Mr. Nopola concludes that the 1987 and 1989 data with the additional 2020/2021 data are suitable for use in this report.

12.7Summary Statement

Mr. Gustin has undertaken extensive verification of the historical data, and he has also reviewed the results from similar verification efforts completed by Integra. This work has identified very few errors in the transcription of field and assay data into the historical mine-site drill-hole databases. In addition, the documentation of gold and silver analytical methods for each historical sample interval allowed Mr. Gustin to identify and remove historical assay data that has insufficient quality for use in the estimations of the current resources.

Explicit modeling of the gold and silver mineralization was the most critical component to the estimation of the project mineral resources. This ‘hands-on’ approach provided meaningful verification of the historical data, whereby Integra infill drill data were found to be consistent with the continuity, widths, and grades of the gold and silver mineralization as defined by the historical drilling. Comparisons of the estimated grades and tonnages of the DeLamar and Florida Mountain areas, with and without substantial input of Integra data, also yielded consistent results and thereby provided further verification of the historical data. Finally, it is important to note that the historical data served as the basis to construct and operate the long-lived and successful historical mining operations at the DeLamar project.

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Integra provided Mr. Gustin lithological and structural interpretations of the DeLamar and Florida Mountain areas. Integra’s geological modeling has been continually refined since the project was acquired, with interpretations of the DeLamar area geology in particular evolving as more drilling, particularly core drilling, was completed. Integra’s geological interpretations have been used by Mr. Gustin as the basis for each successive estimation of the project gold and silver resources. Based on the resource models Mr. Gustin has completed during Integra’s involvement in the project, all of which entail detailed, explicit modelling methods that were completed within the overall context of Integra’s geological interpretations, Mr. Gustin believes the Integra geological models are of sufficient quality to support the current resource modeling.

The authors of this section of this report experienced no limitations with respect to data verification activities related to the DeLamar project. In consideration of the information summarized in this and other sections of this report, the authors believe that the project data are acceptable as used in this report.

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13.0MINERAL PROCESSING AND METALLURGICAL TESTING

This section, prepared under the supervision of Mr. Jack S. McPartland, Senior Metallurgist with McClelland, summarizes the metallurgical testing conducted on samples from the DeLamar and Florida Mountain areas, and historical processing of materials from the same two areas. Estimates of recovery and reagent consumption developed for the processing methods selected for the PFS are included. Mr. McPartland has reviewed the information in this section and believes it is a reasonable summary of the mineral processing, metal recoveries, and metallurgical testing for the DeLamar project as presently understood. The terms “ore” and “whole-ore” used in this section refer to material tested or to be potentially processed and do not imply economic material.

Metallurgical testing is considered in two parts, historical (pre-1990) testing and current (2018-2021) testing for Integra. The current metallurgical testing forms the basis for the recovery and reagent estimates used in this study. The samples used for the current testing generally come from the current mineral resources and are believed to be reasonably representative of the material considered for processing in the PFS. A review of pre-1990 metallurgical testing and processing is presented and contributes to an understanding of the mineralogy and metallurgy for the project. As records of the sample sources for the historical work are incomplete, and in some or most cases the material represented by those samples was likely processed during earlier commercial operations, data from the historical testing are of more limited use than data from the current (2018-2021) metallurgical testing. In general, the oxidation class of the samples used for historical testing (oxide, mixed or non-oxide) was not identified.

Nearly all the historical metallurgical tests and processing data summarized below were originally reported in Imperial units, but in some cases metric weights were reported that were mixed with Imperial distance and concentration units. Use of the original reported units is retained in parts of this section for historical clarity and to avoid awkwardness; the reader is referred to Section 2.2 for the appropriate conversion factors.

13.1DeLamar Area Production 1977 – 1992

13.1.1Mill Production 1977 – 1992

Useful information with respect to mineral processing of DeLamar area gold-silver mineralization by milling and subsequent cyanide leaching is derived from mill production records from the historical open-pit mining operations from 1977 through to the end of 1992. All ore during this time period was mined from the DeLamar area and was processed by crushing, grinding, and cyanide leaching, followed by precipitation with zinc dust and in-house smelting of the precipitate to produce silver-gold doré. After leaching, the solids were concentrated in a series of five thickening tanks and then pumped to a tailing impoundment. During mine closure the tailing were partially dewatered and capped with layers of clay and soil as part of the mine reclamation program.

The DeLamar area produced 421,300 ounces of gold and about 26 million ounces of silver from 1977 through 1992 from 11.686 million tonnes of ore processed with average mill head grades of 1.17 grams Au/t and 87.1 grams Ag/t (Elkin, 1993). The data from Elkin (1993) indicated mill recoveries during the first 15 years of mine operation averaged 96.2% for gold and 79.5% for silver. It should be noted that Elkin (1993) surmised that, “Based on historical records and laboratory testing, the metallurgical recovery of gold is projected to be about 94 percent and 77 percent for silver.”

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13.1.2Cyanide Heap Leaching 1987 – 1990

NERCO constructed a trial cyanide heap-leach pad, which was in operation for the last quarter of 1987 until the final quarter of 1990, using low-grade ROM material dumped by truck and ripped to provide permeability. The material size was reported to be approximately 70% at >20 centimeters (>8 inch).

Mine records from Integra, 2017 indicated that 2,344,037 tons of material, with an average grade of 31.78 g/tonne Ag and 0.41 g/tonne Au were stacked on the heap. It was reported by Integra that the pad base and subsequent stacked material became unstable and began to collapse in mid-1990. Quarterly production records indicate no material was placed on the heap after the second quarter of 1990. In early 1991, the entire heap was removed and placed into the tailing facility. Estimated recoveries were reported to be relatively poor (41% Au and 8% Ag). The incomplete leaching that likely resulted from the pad failure would have adversely affected the reported heap-leach recoveries. These recoveries are not believed to be indicative of recoveries expected from heap leaching of DeLamar oxide and mixed materials.

13.2Historical Testing 1971 – 1989

Multiple metallurgical testing programs were conducted during the 1970s and 1980s on samples from the DeLamar and Florida Mountain deposits. These studies were commissioned by Earth Resources (1970s) and NERCO (1980s). The Earth Resources studies conducted during the 1970s were focused on milling and whole ore agitated cyanidation leaching of the various material types. The NERCO work in the 1980s was more focused on heap leaching of the various material types. A brief summary of those studies follows.

13.2.1Mineralogy from Historical Metallurgical Studies

As reported by Perry (1971), Hazen Research Inc. (“Hazen”) in Golden, Colorado undertook a detailed petrographic and mineralogical study of four sections of drill core from the DeLamar area in 1971. The host rock was described as highly altered porphyritic rhyolite, initially altered to sericite and kaolinite and then silicified and cut by numerous quartz veinlets. Naumannite (Ag2Se) was identified as the chief primary silver-bearing phase, accounting for 75-78% of the total silver present. Argentite (Ag2S) was the other primary silver mineral, accounting for 15-20% of the total silver. Minor gold was found in quartz gangue and as intergrowths in naumannite. Secondary silver minerals in the oxidized portions of core included the silver halide cerargyrite (AgCl), native silver, and argentojarosite, but these minerals together account for only a small fraction of the total silver in the samples.

In 1978, a mineralogical study was conducted by Newmont Exploration Limited (Ahlrichs, 1978) to evaluate improved silver recoveries from Glen Silver feed to the DeLamar processing plant. Samples of feed and plant tailing were evaluated by assay, microscopic examination, XRF, XRD and electron microprobe analysis. There was no information concerning the oxidation classification of the material studied (oxide, mixed or non-oxide). The leached tailing contained approximately 0.5 oz Ag/ton. It was observed that about 90% of the silver was in the form of naumannite (Ag2Se). Remaining silver was present as acanthite (Ag2S). Very little of the silver contained in the tailing was free. Greater than 85% of the unrecovered silver in the tailing occurred as fine inclusions (1 – 26 microns) in quartz. It was concluded that no significant improvement in silver recovery would be gained by finer grinding.

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In 1982, a mineralogy study was conducted by Hazen on eight drill core interval samples from four Glen Silver drill holes (designated GS-1, GS-2, GS-3 and GS-4), to evaluate the mode of occurrence of silver in the samples. The samples were reported to be material from oxidized zones (6 samples), a partially oxidized zone (1 sample) and a non-oxide zone (1 sample). Silver deportment was evaluated using optical and scanning electron microscopy (“SEM”) analysis. The silver mineralization was identified in the oxide samples as silver sulfide, probably acanthite, containing 90% silver and 10% sulfur. The acanthite was associated, mostly as inclusions, with goethite, which was derived from the oxidation of pyrite. Silver minerals were not clearly identified in either the partially oxidized or non-oxide samples, and it was speculated that the contained silver might be present in solid solution with pyrite. Considering the fine grain size of the acanthite and its common occurrence as inclusions in goethite, it was speculated that the material represented by the samples might be refractory to cyanidation, with respect to silver recovery.

13.2.21970s Earth Resources – Hazen Testwork

A series of metallurgical testing programs was conducted on drill hole samples from the DeLamar Sommercamp and North DeLamar areas in the 1970s. This testing included four separate programs conducted by Hazen Research, from 1971 through 1978. The work was conducted on drill composites described as being from the DeLamar, North DeLamar and Sommercamp areas of the DeLamar deposit. Testing included mineralogy, bench scale flotation, agitated cyanidation, salt-roast/cyanidation, ball mill work index testing and solid-liquid separation testing. Test results generally indicated that the samples were amenable to processing by either flotation or agitated cyanidation and that recoveries tended to be higher by cyanidation than flotation.

13.2.3Nerco Minerals Heap Leach Study 1986

A heap-leach study was conducted by Nerco Minerals at an unknown location. Samples tested were identified as being “ore” samples from the Glen Silver and North DeLamar areas and dump samples from the “Sommer Camp” area. Testing included column tests at a -2 inch and 1 inch feed sizes and bottle-roll tests at a -65 mesh feed size. Test results generally showed good amenability of the samples tested to either simulated heap leaching (column testing) at -2 inch or 1 inch feed sizes. Gold recoveries were relatively high (54% – 72%) compared to silver recoveries (12% – 42%). Agitated leach gold and silver recoveries (-65 mesh feed size) were significantly higher.

13.2.4Sullivan Gulch Testing for NERCO 1989

A metallurgical investigation was conducted for NERCO by Hazen on a sample of Sullivan Gulch material. Testing included mineralogical characterization, and evaluation of gravity, flotation, cyanide leaching (both whole feed gravity/flotation products) and pressure oxidation treatment of flotation concentrate.

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The study concluded that multiple processing routes had merit for treating the Sullivan Gulch material, but that no single processing technique was effective for attaining 90% recovery of both gold and silver. Hazen found that gold and silver extractions in excess of 90% could be achieved on the non-oxide samples from Sullivan Gulch using a combination of gravity separation, followed by additional grinding and a second stage of flotation followed by agitation cyanide-leach on the gravity tails (Rak et al., 1989). This combined recovery does not allow for gold and silver losses that would be incurred during subsequent processing of the gravity and flotation concentrates for gold and silver recovery.

13.2.51980s Florida Mountain Testing for NERCO

During the 1980s NERCO conducted column-leach tests using mineralized material from Florida Mountain. No information concerning the oxidation state of the samples tested was available. Statter (1989) summarized metallurgical test work conducted at the DeLamar mine laboratory with Florida Mountain mineralized material in an internal NERCO report. Column-leach and agitation-leach results were reported for Sullivan drill core, Stone Cabin core, and Clark core. In this case, Sullivan core refers to drill core from the Sullivan claim at Florida Mountain, not the Sullivan Gulch area. Results showed generally high column test gold extractions (78% – 93%) with 3 of the 17 column tests showing lower (39% – 52%) gold extraction. Column test silver extractions ranged from 31% to 54%.

Also, Statter (1989) reported a pilot column-leach test was performed in 1988 or 1989 using 14,850 pounds of Stone Cabin “run of dump” material. The test was likely conducted at the DeLamar mine laboratory. Leaching was conducted for 63 days resulting in 15.8% silver recovery and 72.2% gold recovery (Statter, 1989).

13.3Integra 2018-2021 Metallurgical Testing

A multi-phase metallurgical testing program was initiated at McClelland Laboratories, Inc. (“McClelland”) in Sparks, NV by Integra in September 2018, with the primary objectives of evaluating and optimizing processing options for the various material types from both the DeLamar and Florida Mountain deposits. That testing was conducted in two major phases, with testing done in support of the 2019 PEA (referred to as PEA testing), followed by testing in support of this PFS (referred to as PFS testing) conducted from mid-2019 (after the PEA testing) through 2021. Results from PEA testing were summarized in multiple reports (McPartland, 2019a, 2019b and 2019c). Results from the PFS testing were discussed in multiple reports (McPartland, 2020a, 2020b, 2020c, 2021a, 2021b, 2021c, 2021d and Wickens 2020), which were summarized in a combined report (McPartland, 2022).

McClelland maintains ISO accreditation (ISO/IEC 17025:2005) for analytical services, including fire assay, geochemical assay, carbon and sulfur analyses and solution analyses (including gold and silver) presented in this section of the report. All solution analyses, fire assays and geochemical assays conducted as part of the 2018-2019 McClelland metallurgical testing program were conducted following generally accepted industry practices related to quality control and quality assurance, including the use of blanks and standards in each analytical batch.

Metallurgical testing was focused on four main areas: 1) heap leaching of DeLamar oxide and mixed materials, 2) mill processing of DeLamar non-oxide materials, 3) heap leaching of Florida Mountain oxide and mixed materials and 4) mill processing of Florida Mountain non-oxide material. Testing on the oxide and mixed materials was focused on either 2-stage or 3-stage crushing, followed by conventional cyanide heap leaching. Mill processing of the non-oxide material was focused on production of a flotation concentrate from ground feed, followed by regrinding and cyanide leaching of the concentrate, for both deposits.

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The samples used for testing generally were selected to represent the current mineral resources and reserves, and information from this testing campaign forms the primary basis for recovery process selection, and estimates of metal recovery and reagent consumptions, for materials from the DeLamar and Florida Mountain mineral resources. Samples tested have mostly been drill core composites (249), with lesser numbers of RC composites (94, primarily from DeLamar – Sullivan Gulch, non-oxide material) and bulk samples (4).

Samples are identified by oxidation classification, which is based on drill hole logging by the Integra geology staff. The names of two of the three oxidation classifications were changed after the PEA was completed. The three major oxidation classes are now identified as oxide (unchanged, previously called oxide), mixed (previously called transitional) and non-oxide (previously called un-oxide). Where samples used for testing were identified by the old oxidation classification names in the referenced reports, those are changed to the new names herein for clarity purposes. These were changes in naming convention only and did not result in otherwise reclassifying the oxidation class for individual composites.

Drill composites were prepared considering area, oxidation, depth, lithology, alteration, grade and grade continuity. In general, variability composites were at least 6.1 meters (20 feet) in length. Larger composites (typically for column testing or detailed mill testing) generally included 3.0 meter (10 feet) of drill sample below cutoff grade as dilution on either end of the composited interval, as well as below cutoff grade material from within the composited interval. Interval cutoff grades generally used for metallurgical compositing were 0.2 g Au equivalent/tonne for oxide and mixed composites and 0.3 g Au equivalent/tonne for mill (non-oxide) composites.

13.3.1Integra DeLamar Testing

Testing at McClelland on the oxide and mixed material types included column-leach testing at an 80% -12.7mm (0.5-inch) feed size on a total of 13 drill core composites and four bulk samples. The four bulk samples were also each column tested at two coarser feed sizes (100% -200mm and 80% -50mm) (8 inches and 2 inches). Detailed head analyses and a comparative bottle-roll test (80% -1.7mm feed size) (10 mesh) were conducted on each sample. In addition, a total of 57 variability drill composites were prepared for the same bottle-roll testing and head analyses. Load/permeability (hydraulic conductivity) testing was conducted on select column residues to evaluate permeability under simulated commercial heap stack heights.

Testing on non-oxide material included evaluation of gravity concentration, whole ore grind/leach and flotation with flotation concentrate processing by either regrind/cyanidation, roast/cyanidation or Albion oxidation/cyanidation treatment methods. Gravity concentration testing and most of the flotation optimization work was done during the PEA testing. Flotation concentrate processing testwork was done mostly during the PFS testing. Comparative heap-leach variability bottle-roll tests (80% -1.7mm feed size) were also conducted on select non-oxide composites.

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13.3.1.1DeLamar Samples

A summary of the samples comprising the metallurgical composites is shown in Table 13.1. Composites were prepared from a total of 50 drill holes, 34 of which were from 2018 or 2019 drilling and 16 of which were from 2020 drilling. Of the 2018 drill holes, 11 were from RC drilling. These RC holes mostly were drilled in the Sullivan Gulch area. The remaining 2018 and all the 2019 and 2020 samples were drill core. In general, the composites from the 2018/2019 drilling were used for the PEA testing.

Table 13.1 Drill Hole Composite Summary, DeLamar PFS Metallurgical Testing

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* Reverse circulation (RC) drill hole.

1) Samples were composited based on oxidation, lithology, alteration, grade and continuity. In some cases, not all drill sample within the noted range was used for composites.

2) Composite contains material from multiple oxidation classes.

3) Part of an oxide master composite drawing from more than one drill hole (3 oxide master composites total).

4) Part of a mixed master composite drawing from more than one drill hole (2 mixed master composites total).

5) Part of a non-oxide master composite drawing from more than one drill hole (4 non-oxide master composite total).

Note: The multi-hole “master composites” include 3 oxide, 2 mixed and 4 non-oxide composites.

Most of the PFS testing was conducted on composites from the 2020 drilling, though some 2018/2019 drilling was also used for these composites, mainly for non-oxide composites. A map showing the drill hole collars for the DeLamar materials used for the Integra metallurgical testing are shown in Appendix B Figure 1.

The four bulk samples submitted for testing each weighed about three tonnes and were approximately -350mm (-14-inch) in size. These samples were excavated from the DeLamar (Glen Silver) area in March of 2019 for heap-leach testing (Jordan, 2019). The bulk samples included one oxide material sample (sample 4307-B) and three mixed material samples (samples 4307-A, C and D). The mixed material samples were described as either “Trans Clay” (4307-A) or “Trans Hard” (4307-C and D). The “Trans Clay” sample was selected to represent material with an elevated clay content, determined visually.

Detailed head analyses conducted on each of the metallurgical composites and bulk samples included fire assay to determine gold and silver content, cyanide shake analysis to determine cyanide soluble gold and silver, carbon and sulfur speciation analyses and a multi-element ICP scan. Head assays showed the composites contained between 0.09 and 12.7 g Au/t (0.74 g Au/t average) and between 3 and 491 g Ag/t (39 g Ag/t average). Average gold head grades for the oxide, mixed and non-oxide composites were 0.45, 0.56 and 0.91 g Au/t, respectively. Respective average silver head grades were 22, 29 and 50 g Ag/t.

Cyanide soluble to fire assay (“CN/FA”) gold ratios for the oxide, mixed and non-oxide composites averaged 83.0%, 60.5% and 31.6%, respectively. Silver CN/FA values for the oxide, mixed and non-oxide composites averaged 61.7%, 53.5% and 34.7%, respectively.

Average sulfide sulfur head grades for the oxide, mixed and non-oxide composites were 0.14%, 0.61% and 2.12%, respectively. Respective average sulfate sulfur levels were 0.17%, 0.56% and 0.91%. Organic (non-carbonate) carbon content generally was negligible (<0.1%) and no indications of preg-robbing were observed during any of the metallurgical testing.

No strong correlation between sulfide grade and CN/FA ratio was noted for gold or silver. In general, the oxide composites contain low levels of sulfide sulfur and display relatively high gold CN/FA ratios. The mixed composites generally contain <1.5% sulfide sulfur and display highly variable gold CN/FA ratios. Mixed composites containing less than about 0.3% to 0.5% sulfide sulfur generally gave elevated (>50%) CN/FA ratios. Trends were similar when considered by area. The non-oxide composites generally contained >1.0% sulfide sulfur and display highly variable, but generally lower, gold CN/FA ratios.

The DeLamar composites contained low to negligible concentrations of potentially deleterious elements, other than sulfur. Only 10 of the 180 composites assayed for copper contained greater than 50 ppm Cu. Average copper grades were 13 ppm for the oxide and 22 ppm for both the mixed and non-oxide composites. Average mercury head grades were 0.58 ppm (oxide), 0.90 ppm (mix) and 0.47ppm (non-oxide). Respective average selenium concentrations were 18.5, 29.2 and 35.6 ppm for the three oxidation classes. Selenium concentration tended to increase with increasing silver concentration, which would be consistent with the known presence of the silver selenide mineral naumannite (Ag2Se).

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13.3.1.2DeLamar Mineralogy

DeLamar Bulk Mineralogy/Textural Analysis

A total of 67 metallurgical samples from the McClelland PEA and PFS testing program were submitted to Vidence, Inc., in Burnaby, British Columbia for automated SEM scans to determine the mineralogy and texture of the materials. The results were summarized by Vidence (Enter, 2021). The samples included 47 from the DeLamar deposit, of which 24 were oxide, 10 were mixed and 13 were non-oxide material type samples. The remaining samples were from the Florida Mountain deposit (discussed in Section 13.3.2).

The main mineralization identified in the DeLamar deposit across all oxidation states was quartz rich, up to 95%, with feldspar ranging from absent to 23%. Plagioclase was not identified in any of the samples analyzed for the deposit. Pyrite was the primary sulfide mineral identified and was present in trace amounts in all areas of the mixed mineralization except for the DeLamar North region. Pyrite was also detected in trace amounts in the Glen Silver and Sommercamp oxide mineralization as well. Non-oxide mineralization was analyzed from Sullivan Gulch and Sommercamp. Pyrite was the main sulfide mineral with abundances up to 5.9%. Sphalerite, arsenopyrite, and chalcopyrite were also present in trace amounts. Alunite and jarosite were identified as sulfate phases present in the Glen Silver, South Wahl, and Ohio oxidized and mixed mineralization with abundances up to 3.1%.

Clay species were common throughout all the areas of the DeLamar deposit with varying amounts of muscovite, illite, and kaolinite identified. The Milestone samples exhibited significantly more total clay in the oxidized mineralization (30%) than the samples of mixed (3%) that were analyzed. The clays within the Milestone area exhibit signs of being potentially sensitive to water. South Wahl had the most intensive clay study with 15 samples analyzed. Samples exhibited an inverse relationship between illite and kaolinite concentration. Low kaolinite content was also correlated with high abundance of K-feldspar. The South Wahl oxide samples generally contained relatively high levels of clays, with illite content varying from 1.78% to 40.76% and kaolinite content varying from 4.11% to 40.76%. The South Wahl mixed material exhibited lower levels of kaolinite with a similar range of illite content. Further work would be required to confirm potential water sensitivity/swelling tendencies of the clays for the entirety of the DeLamar deposit.

DeLamar Non-Oxide Sample Gold and Silver Deportment Analysis

A mineralogical assessment was conducted by BV Minerals, in Richmond, British Columbia on five DeLamar non-oxide (identified during testing as non-oxidized) composites, including two from Glen Silver (composites designated 4307-192 and 4307-199) and three from Sullivan Gulch (composites 4307-012, 120 and 165), as part of the McClelland PFS testing program, to determine general mineralogy as well as gold and silver deportment, liberation characteristics and mineral associations. The results were summarized in BV Minerals (2020). Gold CN/FA ratios for these composites generally were low (12.6% – 37.1%), with the exception of Sullivan Gulch composite 4307-120 (Au CN/FA of 66.4%). Testing included QEMSCAN particle mineral analysis (“PMA”) and trace mineral search (“TMS”). Gold deportment relied heavily on analysis of gravity concentrate produced from each composite because of the low (gold) grade nature of the samples.

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The samples were described as presenting low degrees of sulfide mineralization (2.6% – 9.8% sulfide minerals). Pyrite/marcasite was the primary sulfide mineral. The composites also contained 0.1% to 0.7% arsenopyrite. Trace amounts of other sulfide minerals were also noted. Non-sulfide minerals included quartz, muscovite/illite, K-feldspar and kaolinite (clay), in order of mineral abundance.

All of the gold found in the five composites presented as native gold or electrum. Gold grain sizes ranged from 0.5 microns to 50 microns and averaged 0.8 microns to 11.8 microns in circular diameter. The majority of gold occurrences were <10 microns. Over half of the gold in 4307-012 (SG), 165 (SG) and 199 (GS) was finer than 2 microns and was mostly locked with pyrite and non-sulfide gangue. Relatively coarse (20 - 50 microns) gold was also observed in two of the Sullivan Gulch composites (4307-012 and 4307-120). These coarse grains are expected to be gravity recoverable. It was noted by the mineralogist that based on comparison between the grade of the samples evaluated and the amount of gold identified during the examination, it was possible to conclude that the samples also likely contain significant quantities of gold in solid solution with sulfide minerals ("invisible" refractory gold). Only 5% to 30% of the gold observed in the other three composites presented with exposed surfaces. Some difficulties in recovering gold from these composites was anticipated because of the relatively fine-grained gold observed (0.8 microns to 5.7 microns average).

Silver was present in the samples mainly as naumannite (Ag2Se), pyrargyrite (Ag3SbS3), freibergite ((Ag,Cu,Fe)12(Sb,As)4S13) and, in the case of the Glen Silver composites, canfieldite (Ag8SnS6). The relative abundance of these minerals varied significantly between composites. Silver grain sizes ranged from 0.5 microns to 20 microns and was most commonly present in the 10 microns to 20 microns size range.

Silver liberation ranged from about 9% to 38% at the size studied (nominal 90 microns). Unliberated silver minerals were predominantly locked in pyrite and non-sulfide gangue. Between 50% and 95% of the silver in all composites but 4307-192 (GS) presented exposed surfaces. Silver recoveries greater than 50% were expected by leaching.

Diagnostic leach tests (“DLTs”) were conducted at McClelland (McPartland, 2020a) as part of the same study, but on five different non-oxide composites and composite 4307-012 (also submitted to BV Minerals for mineralogy). These composites included two from Glen Silver (4307-190 and 4307-191) and three from Sullivan Gulch (4307-012, 4307-029 and 4307-149/153).

A total of four sequential leach steps were performed as part of the DLTs, including direct cyanidation, hydrochloric acid digestion followed by cyanidation, aqua regia digestion followed by cyanidation, and roasting followed by cyanidation of the calcine. The final leached residue was fire assayed. These procedures are designed to determine gold deportment from the samples.

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Testing confirmed that most of the gold losses to grind/leach (cyanidation) tailing resulted from locking of contained gold values in difficult to oxidize sulfide mineral particles (likely pyrite/arsenopyrite). Test results indicated that oxidative pretreatment (such as POX, roasting or Albion) on either ore or flotation concentrate should be effective in substantially improving gold recovery from the non-oxide materials. Gold deportment as determined by the diagnostic leach test procedures generally was consistent with observations made during the BV Minerals mineralogical characterization of other DeLamar non-oxide samples. Only composite 4307-012 (Sullivan Gulch) was used for both mineralogy and diagnostic leach testing.

Cyanide extractable gold for the Sullivan Gulch composites generally ranged from 25.1% to 51.6%. Gold recovery by cyanidation of composite 4307-047 was higher (78.5%). For the three other SG composites, most of the gold reporting to the cyanidation tailing was indicated to be locked in difficult to oxidize sulfide minerals, such as arsenopyrite or pyrite, as indicated by the relatively large portion (37.1% – 61.8%) of the total contained gold that was recovered after aqua-regia digestion. Small amounts (7% – 8%) of the contained gold was recovered after subsequent roasting, but this was attributed to incomplete sulfide mineral leaching during the aqua regia digestion stage. Less than 3% of the gold contained in the Sullivan Gulch non-oxide composites reported as being locked in silica, as indicated by fire assay of the final test residue.

The two Glen Silver composites tested were more refractory to cyanidation treatment (gold recovery of 7% – 11%) but followed similar trends with respect to the gold losses to the cyanidation tailing. A very large portion (84%) of the total contained gold was indicated to be locked in difficult to oxidize sulfide minerals such as pyrite and arsenopyrite. Small amounts of gold were indicated to be associated with either carbonates, iron oxides or more reactive sulfides and with silicates. Again, the relatively large amount of gold locked in difficult to oxidize sulfide minerals will likely require oxidative pretreatment for liberation and recovery by subsequent cyanidation.

13.3.1.3DeLamar Heap-Leach Testing

A detailed heap-leach testing program was conducted in two phases on oxide and mixed samples from the DeLamar deposit. Those phases consisted of the PEA work conducted on 2018 and 2019 drill samples and four bulk samples (McPartland, 2019; 2020b; Wickens, 2020) and on the PFS work conducted on 2020 drill samples McPartland, 2022). For the ease of discussion, results from both programs are presented together here where appropriate.

Both phases of testing included bottle-roll tests on drill hole variability composites (58 total). These composites generally were comprised of 3-meter (9.8-foot) to 10-meter (32.8-foot) continuous runs of drill core from a single hole (8.5-meter average; 25 drill holes total) (27.9-foot) and included 41 oxide and 17 mixed type composites. Column-leach tests were also conducted on a total of 17 samples (four bulk and 13 drill core composites). Tests were conducted at an 80% -12.7 mm (0.5 inch) feed size on all but one sample. Comparative tests were conducted on the four bulk samples and on three of the core composites at a minus 50 mm (2-inch) feed size. A single core composite was tested only at the minus 50 mm feed size. Tests were conducted in a manner to determine gold and silver recoveries, recovery rates and reagent requirements, under simulated heap leaching conditions. A comparative bottle-roll test was conducted on each of the column test samples, at the same size as used for the variability tests (80% -1.7 mm) to help establish the relationship between bottle-roll and column test recoveries.

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Detailed head analyses, including fire assay, cyanide solubility analysis, carbon and sulfur speciation analysis and multi-element ICP scans were conducted on each sample. Oxide composite gold head grades ranged from 0.09 to 2.84 g Au/t and averaged 0.45 g Au/t. Oxide composite silver head grades ranged from 3 to 69 g Ag/t and averaged 22 g Ag/t. Sulfide sulfur grade for the oxide composites generally was below 0.15% and averaged 0.13%. CN/FA ratio for gold generally was high and averaged 83%. In the three cases where oxide composites gave anomalously low CN/FA ratios (<50%), bottle-roll test gold recoveries were substantially higher.

Mixed composite gold head grades (labeled transitional for the PEA testing) ranged from 0.16 to 1.82 g Au/t and averaged 0.56 g Au/t. Mixed composite silver head grades ranged from 4 to 96 g Ag/t and averaged 29 g Ag/t. Sulfide sulfur grade for the mixed composites were variable (0.03% – 1.63%) and averaged 0.61%. CN/FA ratio for gold ranged from 28.5% to 100% and averaged 61%.

Oxide composite CN/FA ratio was not correlated to sulfide head grade, but mixed composite CN/FA ratio was to a moderate degree. Mixed composites containing <0.3% sulfide sulfur generally displayed elevated (>70%) CN/FA ratios for gold. Mixed composites containing >0.5% sulfide generally displayed relatively low (<50%) CN/FA ratios.

DeLamar Heap-Leach Variability Bottle-Roll Testing

Bottle roll tests were conducted on each of the oxide and mixed variability composites, at an 80% -1.7 mm (10 mesh) feed size, to evaluate potential for heap leaching and material variability. A summary of bottle-roll test conditions and results for the oxide and mixed variability composites are presented in Appendix B Table 1 and Appendix B Table 2, respectively, and are presented graphically in Figure 13.1.

Figure 13.1 Gold Recovery, Bottle-Roll Tests, DeLamar PFS Variability Composites

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Variability bottle-roll test results showed that, in general, the oxide composites were readily amenable to agitated cyanidation treatment, at the 1.7 mm feed size. Gold recoveries from the 41 composites ranged from 55.3% to 96.3% and averaged 82.1% in four days of leaching. Only three of the 41 composites tested gave gold recoveries of <65%.

Silver recoveries obtained from the oxide composites were more variable and generally lower, compared to the gold recoveries. Silver recovery from the 41 composites ranged from 14.6% to 76.5% and averaged 39.2%, in four days of leaching. About one-half of the oxide composites gave silver recoveries of <40%.

A much smaller number (17) of mixed and blended (containing more than one oxidation type) composites were tested. Gold recoveries obtained in four days of leaching ranged from 26.5% to 83.3% and averaged 54.1%. Eight of 17 of the composites gave gold recoveries of <50%. Gold recoveries were not strongly correlated to composite sulfide grade, but in the case of the mixed material were moderately well correlated to CN/FA ratio. This correlation will be useful for further development of the heap-leach recovery models for the DeLamar mixed materials.

Silver recoveries were more varied and on average were about 8% higher than obtained from the oxide composites. Silver recoveries from the mixed composites ranged from 20.0% to 87.7% and averaged 47.3%. About one-third (6 of 17) of the mixed composites gave silver recoveries less than 40%.

Reagent consumptions were low. Cyanide consumptions for the oxide and mixed composites averaged 0.15 and 0.19 kg NaCN/t. Respective lime requirements averaged 2.2 and 1.8 kg/t of ore. Only three of the 61 composites had cyanide consumption greater than 0.3 kg NaCN/t. Only seven of the composites had a lime requirement greater than 3.0 kg/t.

Variability bottle-roll tests were also conducted on a total of 90 non-oxide drill composites, at an 80% -1.7 mm (10 mesh) feed size. These composites were primarily from the Sullivan Gulch area (74 samples), with 16 from the other DeLamar areas. Procedures were the same as described above for the oxide and mixed composites.

Bottle roll testing at the 1.7 mm feed size generally indicated that the non-oxide material will not be amenable to heap-leach processing. Gold and silver recoveries were low. Average gold recoveries from DeLamar (non-Sullivan Gulch) composites and from Sullivan Gulch composites were 16.8% and 32.1%, respectively. Average respective silver recoveries were 25.0% and 36.4%. Only five of the composites, all from Sullivan Gulch, gave gold recoveries of >50%.

DeLamar Column-Leach Testing

Column leach tests were conducted on each of 12 column test drill core composites and four bulk samples, representing DeLamar oxide and mixed material types, at an 80% -12.7 mm (0.5 inch) feed size, using 10 cm (4 inch) or 15 cm (6 inch) diameter by 3-meter (9.8-foot) high leaching columns. A 9.6 Lph/m2 solution application rate and a cyanide concentration of 1.0 g NaCN/L were used. Column test duration generally was about 60 days. The core composite feeds were not agglomerated. Lime was added for pH control to the dry test charges before leaching. The bulk sample 80% -12.7 mm charges were agglomerated before column leaching using cement (5.0 – 10.0 kg/t). Comparative column-leach tests were conducted on the four bulk samples, at 100% -200 mm (8-inch) and 80% -50 mm (2-inch) feed sizes, to determine feed size sensitivity. Larger diameter columns were used for the -200 mm feeds, (60-cm diameter) (23.6-inch diameter), and 50 mm feeds (25-cm diameter) (9.8-inch diameter). Column leach tests were also initiated at a 100% -50 mm (-2-inch) feed size on three of the same 12 column test core composites tested at 12.7 mm (0.5-inch) and on a fourth column test core composite tested only at the -50 mm feed size. Those -50mm tests have not yet completed and are not included in the results presented here. Comparative bottle-roll tests were conducted on each column test composite, at an 80% -1.7 mm (10 mesh) feed size to establish the relationship between recoveries from the two tests. Summary results from the bulk sample column and bottle-roll tests and size versus recovery curves for the column tests are presented in Table 13.2 and Figure 13.2, respectively.

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Table 13.2 PEA Column-Leach and Bottle-Roll Tests, DeLamar and Glen Silver Bulk Samples

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Figure 13.2 PEA Column Column-Leach Gold and Silver Recovery vs. Feed Size

(Glen Silver Bulk Samples)

Column testing showed that the Glen Silver bulk samples were amenable to heap-leach cyanidation treatment at the feed sizes evaluated. Gold recovery obtained from the oxide sample (4307-B) was not sensitive to feed size and ranged from 86.4% at the -200 mm feed size to 88.0% at the 80% -50 mm (-2-inch) feed size. Gold recovery rate was rapid and increased with decreasing feed size.

The Glen Silver mixed bulk samples were more sensitive to feed size. Gold recovery from sample 4307-A (“trans clay”) increased from 64.2% at the -200 mm (-8-inch) feed size to 73.4% at the 80% -12.7 mm (-0.5-inch) feed size. Gold recovery from mixed sample 4307-C (“trans hard”) improved from 87.5% at the -200 mm feed size to 92.5% at the 12.7 mm feed size. This sample contained only 0.06% sulfide sulfur and may be better classified as oxide material. Gold recovery from mixed sample 4307-D (“trans hard”) improved from 50.0% at the -200 mm feed size to 67.7% at the 12.7 mm feed size. Gold recovery rates were slowest from the coarsest feeds, but generally were fairly rapid for the 50 mm and 12.7 mm feeds. Gold recovery rates from mixed sample 4307-D were very slow, and gold extraction was progressing at a significant rate from these feeds when leaching was ended.

Cyanide consumption was low to moderate (<1.5 kgNaCN/t of ore) for all bulk samples. The lime added to the -200 mm and 50 mm column charges before leaching was sufficient for maintaining leaching pH. The 12.7 mm feeds were agglomerated with cement before leaching, as a precautionary measure.

Core composite column test results are presented in Table 13.3.

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Table 13.3 PFS Column-Leach Test and Bottle-Roll Test Results, DeLamar Core Composites

The oxide composites were all amenable to simulated heap leaching, at an 80% -12.7mm (-0.5-inch) feed size. The mixed composites were more variable with respect to gold and silver recoveries obtained at an 80% -12.7 mm feed size, but generally were amenable to simulated heap leaching. Gold recoveries from the oxide core composites ranged from 67.6% to 93.2% and averaged 78.1%, in 53 to 59 days of leaching.

Column test silver recoveries obtained from the oxide composites ranged from 8.8% to 29.3% and averaged 16.7%. Silver extraction, and in some cases gold extraction, were progressing at a slow rate when leaching was terminated (<60 days). Silver recovery in particular would be expected to increase significantly with longer leaching cycles.

Column test gold recoveries from the mixed composites were lower and variable. Gold recovery obtained from the high sulfide (1.37% S) Glen Silver composite (4522-050) was very low (24.2%). The CN/FA ratio for this composite was also very low (39.4%) and the composite is thought to be better classified as non-oxide material. Column test gold recovery from the other Glen Silver mixed composite was substantially higher (75.9%). These conflicting data demonstrate the need for further refinement of the oxidation logging and modelling for the DeLamar deposit. Column test gold recoveries from the two other mixed ore type composites were 47.8% (Milestone Comp. 4522-046) and 52.4% (South Wahl Comp. 4522-072).

The Milestone column test composite was unique in that it did not contain detectable sulfide sulfur, though the material comprising the composite was mostly (80%) logged as mixed material (20% logged as oxide material). This composite gave a lower-than-expected column test gold recovery (47.8%), considering it contained no sulfide minerals. Cyanide soluble gold content for this composite was 54.2% Au, which was consistent with the column test gold recovery. Variability bottle-roll tests conducted on Milestone composites (Appendix B Table 1, Appendix B Table 2) showed variable but generally high (78.5% average) gold recovery from the oxide composites and variable but generally lower (49.7% average) gold recovery from the mixed composites.

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Column test silver recoveries from the four mixed composites were variable and ranged from 15.8% to 63.6%. As discussed below, silver extraction generally was incomplete when the tests were ended, and silver recovery tended to be higher for the samples with higher sulfide sulfur grades.

Gold and, in particular, silver extractions were progressing at a slow rate when leaching was terminated for the mixed samples. Longer leaching cycles would be expected to improve gold recoveries incrementally and silver recoveries moderately.

In general, neither gold nor silver recovery were strongly correlated to lithology, head grade or any of the trace metal analyses conducted on each of the composites. Except for the Milestone composite (Comp. 4522-046), gold recovery was fairly strongly inversely correlated to sulfide sulfur grade. Without exception, other non-Milestone column test feeds that contained less than 0.5% sulfide sulfur gave gold recoveries of >65%. Gold recovery was lower for samples that contained > 0.8% sulfide sulfur. This trend is consistent with the mineralogical data available from the DeLamar non-oxide material, which shows a significant portion of contained gold locked in sulfide minerals.

Silver recovery exhibits a moderate, positive correlation with sulfide grade. This effect may be muted by the incomplete silver extraction obtained from the columns because of time constraints. This correlation is consistent with the generally higher silver recoveries observed from the mixed material compared to the oxide material. Further testing and mineralogical studies will be required to better understand the variability in results.

Cyanide consumptions were moderate and ranged from 0.91 to 1.54 kg NaCN/t for the oxide composites and from 1.12 to 1.59 kg NaCN/t for the mixed composites. Considering these cyanide consumptions and the consistently low cyanide consumptions observed during bottle-roll testing, it is expected that commercial cyanide consumptions will be substantially lower (<0.5 kg NaCN/t) during commercial production. The 1.1 to 3.3 kg/t lime added before leaching was sufficient for maintaining pH throughout the leaching cycles.

Head and tail screen analyses were conducted on the feed and residue from each column test. Recovery by size data generally indicated for the oxide material that finer crushing or grinding would significantly improve silver recoveries but would have minimal effect on gold recovery. Recovery by size data from the mixed composites, including the Milestone composite, generally indicated that fine grinding would be required to maximize gold and silver recovery by cyanidation.

The correlation between column-leach test recoveries and bottle-roll test gold recovery is strong and should be useful for predicting heap-leach recoveries from bottle-roll test data. The correlation for silver is less strong but should still be useful.

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None of the DeLamar samples displayed any solution percolation/permeability problems during column leaching. The bulk sample 12.7 mm (0.5-inch) column feeds were agglomerated with cement as a precautionary measure before leaching. Later testing (discussed below) indicated that those feeds did not require agglomeration pretreatment. None of the other DeLamar column feeds were agglomerated.

DeLamar Load/Permeability Testing

Select 12.7mm (0.5-inch) feed size column-leached residues from the PFS column testing program were submitted to Geo-Logic Associates, in Sparks, Nevada, for load/permeability tests (test procedure USBR 5600-89). The residues selected were those leached in 15-cm (6-inch) diameter columns which provided sufficient sample for this testing, along with tail screen analyses. The tests were conducted to evaluate permeability characteristics (hydraulic conductivity) under expected commercial heap stack height compressive loads (up to 150 meters) (492 feet). Results from the load/permeability tests showed that, without agglomeration pretreatment, marginal to poor permeabilities were obtained at simulated heap stack heights greater than 30 to 80 meters (98.4 to 262.5 feet). In four cases (Comps. 4522-046, 052, 053 and 072) permeabilities below the equivalent expected heap irrigation rate (6.0 Lph/m2) would be expected at greater than about a 10 to 50-meter (32.8 to 164-foot) heap stack height. These results indicate that agglomeration pretreatment will be required for heap leaching of these 12.7 mm feeds.

After these results were reviewed, a weighted average composite of the 12.7 mm column residues was prepared for agglomeration testing. A series of agglomerate strength and stability (“plunk and dunk”) tests were conducted on that composite to optimize cement binder addition. This test procedure consists of preparing small quantities of agglomerated ore (1.0 kg) (2.205 pound) per binder addition tested with varied cement additions and subjecting those agglomerates to stability and strength testing to evaluate the durability of the agglomerates at the cement doses tested. Tests ranged from 1.5 to 8.0 kg/t with lime, plus cement also evaluated at lower binder additions. Results indicated a cement addition of approximately 3 to 5 kg/t was optimal.

Based on results from the agglomeration testing, two agglomerated samples (using 3.0 kg/t and 5.0 kg/t cement) were prepared from the same column residue composite at McClelland and submitted to Geo-Logic for load/permeability testing. Results showed that the agglomerates prepared using 5.0 kg/t performed well and permeabilities were high at simulated heap stack heights of up to 137 meters (449.5 feet) (the highest load tested). Permeability for the 3.0 kg/t cement agglomerates also performed well and acceptable permeabilities were observed at simulated heap stack heights of up to about 80 meters (262.5 feet). These results demonstrate that, for the material types displaying poor permeability characteristics, agglomeration using 3.0 kg/t cement should be sufficient to ensure adequate solution percolation characteristics for heap stack heights of up to about 80 meters (262.5 feet). For higher heap stack heights, a higher cement binder addition of 4.0 to 5.0 kg/t will likely be required.

Load/permeability tests were also conducted on the four bulk samples from the PEA testing at the 12.7 mm feed size. Fresh material was used for these tests. Results showed acceptable permeability under load without agglomeration pretreatment.

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13.3.1.4Integra DeLamar Mill PEA Testing

Testing conducted at McClelland during the 2019 PEA study (Gustin et. al., 2019) included evaluation of grind-leach, gravity concentration and flotation processing of the DeLamar non-oxide materials. Testing to optimize flotation processing was also conducted.

DeLamar Agitated Cyanide Leach Testing

Thirty non-oxide composites from DeLamar, Glen Silver, Sommercamp and Sullivan Gulch, along with three mixed, one oxide and two blended (mixed/non-ox) composites, were used for a bottle-roll leach test at an 80% -75 microns (200 mesh) feed size, to evaluate amenability to “whole-ore” milling/cyanidation treatment. Tests were conducted using 72-hour leach cycles at 40% solids and a cyanide concentration of 1.0 g NaCN/L solution.

Gold recoveries obtained from the one oxide and three mixed composites ranged from 61.8% to 88.9%, in 96 hours of leaching, indicating good potential for milling/cyanidation treatment. Gold recoveries obtained from the non-oxide and blended (mixed/non-ox) composites, at the 75 microns feed size, were highly variable, ranging from 6.8% to 81.1%. Gold recovery was not correlated to sulfide sulfur content, sample depth (down-hole) or elevation. Silver recoveries also were variable and ranged from 14.3% to 76.7%. While these results indicate the potential for milling/cyanidation treatment for a relatively small portion of the DeLamar non-oxide material, particularly from the Sullivan Gulch area, recoveries were highly variable. The high variability in recoveries makes direct grind-leach of the non-oxide materials an unlikely process option.

DeLamar Gravity Concentration and Flotation Testing

A series of scoping-level gravity concentration tests, with bulk sulfide flotation on the gravity tailing, was conducted on nine non-oxide drill core composites from Sullivan Gulch and Glen Silver. These tests were conducted to obtain preliminary information regarding the effectiveness of upgrading the DeLamar non-oxide material by conventional gravity concentration and flotation processing methods. The samples tested included eight composites from Sullivan Gulch and one composite from Glen Silver. Samples of 1.0 kg were ball milled to 80% -75 microns (200 mesh), subjected to gravity concentration and the resulting gravity rougher tailing were subjected to bulk sulfide flotation treatment.

The Sullivan Gulch composites generally responded well to gravity concentration, followed by bulk sulfide flotation treatment, at an 80% -75 microns feed size. The gravity rougher concentrates contained between 2.7% and 5.6% of the “whole-ore” mass and represented average gold and silver recoveries of 34.9% and 26.4%, respectively. The resulting gravity tailing generally responded well to bulk sulfide flotation treatment. The combined gravity/flotation rougher concentrate produced from six of the eight Sullivan Gulch composites tested was equivalent to an average of 19% of the “whole-ore” weight and contained an average of 89% of the total gold and 91% of the total silver. The remaining two Sullivan Gulch composites also gave high gold and silver recoveries, but the mass pull during flotation was anomalously high at about 40% of the “whole-ore” weight. Carry-over of clay minerals to the concentrates appeared to be responsible for the higher mass pulls observed during these tests. It was noted that an improved response to flotation should be expected through optimization of flotation conditions. Combined gravity-flotation recoveries obtained from the single Glen Silver composite were somewhat lower and were equivalent to 74.2% gold and 78.0% silver.

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Based on results from the preliminary tests described previously, two master composites were prepared from each of the non-oxide Sullivan Gulch and Glen Silver materials. Those composites were used for optimization of flotation conditions. Flotation testing on those composites was conducted without gravity concentration. Parameters evaluated included feed size for all four composites, along with reagent additions and rougher concentrate regrind for the two Sullivan Gulch composites. A total of 24 tests were conducted on the four composites. Representative test results are shown in Appendix B Table 3.

Both Sullivan Gulch composites responded reasonably well to bulk sulfide flotation treatment, with gold and silver recoveries to the flotation rougher concentrates of approximately 90% to 95%. Selectivity for the initial tests was lower than desired, and flotation rougher concentrate mass pulls were generally about 20% to 30%.

Optimization testing was successful in decreasing flotation mass pull to about 9% to 11%, while maintaining gold recovery, either to cleaner or rougher concentrate, at between 86% to over 90%, and silver recovery above 90%. It was noted that gold and silver recoveries in excess of 90% are expected to be obtainable at a concentrate mass pull of about 10% to 13%. Grind optimization test results indicated that a grind size of 80% -150 microns (100 mesh) will likely be sufficient for maximizing gold and silver recoveries by flotation of the Sullivan Gulch non-oxide material. Rougher flotation recoveries appeared to be incrementally lower at a coarser feed size of 80% -212 microns (65 mesh).

Testing on the Glen Silver non-oxide master composites showed that they gave lower flotation recoveries than the Sullivan Gulch material. In general, gold and silver recoveries of about 75% were achieved with flotation rougher concentrate mass pulls of approximately 14% to 19%. Very fine grinding (80% -45 microns) (325 mesh) was evaluated to determine if rougher flotation recoveries could be improved. Those results indicated that finer grinding was not effective for improving recoveries.

Modified diagnostic leach tests were conducted on the flotation rougher tailing generated from the Glen Silver composites, at the -45 microns feed size, to determine gold deportment of values reporting to the flotation tailing. Results indicate the potential for significantly improving flotation recoveries through continued optimization of flotation conditions, and they suggest the recovery of cyanide-soluble gold from the flotation tailing by leaching can be considered. Mineralogical examination of select flotation tailing is planned to better evaluate causes for gold and silver losses to the flotation rougher tailing from the Glen Silver non-oxide material.

13.3.1.5Integra DeLamar Mill PFS Testing

The PFS mill testing was conducted at McClelland in multiple stages, primarily on non-oxide material types, and included further evaluation of flotation variability and optimization along with testing of flotation concentrate processing by regrind-cyanidation, roast-cyanidation, and “Albion” oxidation-cyanidation. Direct agitated cyanide leaching was evaluated at feed sizes ranging from 80% -75 microns (200 mesh) to 80% -10 microns to evaluate gold and silver liberation characteristics and the potential for very fine or ultra-fine regrinding of flotation concentrate followed by agitated cyanidation.

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The report by Wickens (2020) summarized all testing to December 2020 on DeLamar and Florida Mountain samples at McClelland. The PEA metallurgical testing was described as Test Series 1 through 13 in Wickens (2020). PFS testing related to flotation and flotation concentrate processing conducted through December 2020 was described as Test Series 13, 14 and 18 through 21 in Wickens (2020). Those test series designations are mentioned here where appropriate for ease of reference. Mill testing conducted at McClelland in 2021 as part of the PFS metallurgical testing program was described in a separate report (McPartland, 2021d) and in the summary report (McPartland, 2022).

PFS Flotation Testing

Multiple flotation tests were conducted on a total of 26 drill hole composites, as part of the McClelland PFS metallurgical study. These tests were conducted to further evaluate variability in response to flotation treatment and to generate flotation concentrate for cyanidation testing. Summary results for flotation tests conducted during the PFS testing are presented in Appendix B Table 4, and in Figure 13.3 and Figure 13.4.

Figure 13.3 Glen Silver Flotation Recoveries, PFS Composites

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Figure 13.4 Sullivan Gulch Flotation Recoveries, PFS Composites

Flotation testing generally showed that the Glen Silver composites (four), Sommercamp composite (one) and Sullivan Gulch composites (21) responded well to conventional flotation treatment, at either an 80% -150 microns (100 mesh) or 80% -75 microns (200 mesh) feed size. When the same composite was evaluated at both feed sizes, only done for three Glen Silver composites, recoveries to the rougher concentrate were essentially the same at the two sizes.

Flotation gold recoveries that reported to the concentrates produced from the Glen Silver composites generally were equivalent to between 70.5% and 85.3%. Flotation gold recoveries averaged 77.2% at the 150 microns feed size and 78.0% at the 75 microns feed size, for the three composites tested at both feed sizes. Silver recoveries obtained from the Glen Silver composites generally ranged from 69.7% to 89.9%. Flotation silver recoveries averaged 80.4% at the 150 microns feed size and 80.7% at the 75 microns feed size, for the three composites tested at both sizes.

Glen Silver composite 4307-193/194, which was used to generate concentrate for a regrind/cyanidation test, gave an anomalously poor flotation response. Flotation gold and silver recoveries to the rougher concentrate were only 47.3% and 53.8%, respectively. Sulfide sulfur recovery data was not available for this test. The cause for the poor recovery was not well understood but is believed to be anomalous. Sulfide sulfur recovery for the Glen Silver flotation tests ranged from 88.8 to 97.8% and was not sensitive to feed size.

A single Sommercamp composite was evaluated at an 80% -150 microns feed size. The flotation test with a 10.4% mass pull gave rougher concentrate gold and silver recoveries of about 90%. The second test had a lower mass pull and correspondingly lower recoveries.

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A total of 15 Sullivan Gulch composites were used for flotation testing at an 80% -150 microns feed size and nine were used for flotation testing at an 80% -75 microns (200 mesh) feed size. None of the composites were tested at both feed sizes. Samples tested at the -150 microns (100 mesh) feed size included two Sullivan Gulch latite (“Tsgl”) lithology composites and one composite that contained 40% Tsgl lithology material. The Tsgl lithology composites gave significantly poorer flotation response than any of the other samples tested. Mass pulls for this material tended to be high and recoveries tended low and variable where more than one test was conducted. It was noted during these tests that there was a significant carry-over of clay-like gangue material into the rougher concentrates. Analysis of drill hole logging by Integra staff indicated that the Tsgl lithology represents a relatively small portion, 3% by drill hole length, of the Sullivan Gulch non-oxide material above cut-off grade. Although further testing to optimize flotation processing of this material type should be considered, the impact the material will have should be relatively small. High mass pulls were also noted for several other non-Tsgl composites (4307-207, 208 and 211), but flotation recoveries for those composites were relatively high.

The Sullivan Gulch composite 4307-154 gave an anomalously poor gold and silver recovery at a -150 microns feed size during PFS testing to generate concentrate for regrind/leach testing (Test Series 20). Mass pull for that test (5.4%) was lower than optimal. As reported for the PEA testing, flotation gold and silver recoveries of >80% were obtained at the -150 microns size from the same composite, when a more optimal mass pull (8.2%) was taken.

Omitting the Tsgl lithology material and the anomalous composite 4307-154 test, flotation gold recoveries obtained at the -150 microns and -75 microns feed sizes averaged 85.0% and 92.6%, respectively. Silver recoveries obtained averaged 92.6% and 93.6%, respectively. As noted above, none of the composites were tested at both feed sizes and the differences in average recoveries are believed to result mainly from the difference in samples tested.

PFS Flotation and Flotation Concentrate Processing

Drill hole composites (two each) from Sullivan Gulch and Glen Silver were prepared for flotation testing, to confirm flotation response and to generate “high quality” flotation rougher concentrate for testing to evaluate processing techniques for recovery of gold and silver from the flotation rougher concentrate. Tests were conducted on 8 kg lots of sample, milled to 80% -150 microns (100 mesh). Flotation was conducted using procedures optimized during earlier testing. Results from the flotation tests were presented in Appendix B Table 4 (referenced Test Series 13). In general, the flotation gold and silver recoveries in these tests were somewhat lower than previous tests. The rougher concentrate and tailing grades were higher, and mass pulls to concentrate were lower.

Sample splits of 100 g (3.53 ounces) from the Glen Silver and Sullivan Gulch flotation concentrate master composites were combined, finely ground and subjected to mechanically agitated cyanide tests. The concentrate samples were reground to >99% -25 microns before cyanide leaching under relatively intensive conditions. Test conditions and summary results are shown in Appendix B Table 5. Results showed that even after a relatively fine regrind, gold recoveries remained relatively low. Gold recoveries from the Glen Silver and Sullivan Gulch rougher concentrates were 33.9% and 66.5%, respectively, of gold contained in the concentrate. The gold CN/FA ratios for the two respective composites (“whole ore”) were 25.2% and 51.9%. Comparison to the concentrate leach recoveries indicates that the benefit of regrinding on gold recovery was relatively small. In contrast, silver recoveries from the Glen Silver and Sullivan Gulch rougher concentrates were 82.1% and 92.3%, respectively, of silver contained in the concentrate. These silver recoveries were significantly higher than the silver CN/FA ratios (49.5% – 56.0%), indicating a substantial improvement realized by very fine regrinding.

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Splits of the same two rougher concentrate composites, as well as a split from a cleaner concentrate generated from the composite 4307-154 rougher concentrate, were shipped to the Nevada Gold Mines Carlin Metallurgical Lab in Nevada for roaster testing (Hill, 2019). A roaster test with carbon-in-leach (“CIL”) test on the roast calcine was conducted on each sample. .

Compared to the fine grind-cyanide leach tests, the roasting tests on Sullivan Gulch concentrates did not result in a significant increase in gold or silvery recovery. The roasting test on Glen Silver concentrate resulted in increased gold recovery, but the test showed decreased silver recovery. The roaster tests performed at Nevada Gold Mines had mixed results, and it should be noted these tests were not optimized. It is not clear how optimized performance might improve metal recoveries.

For the Glen Silver sample, the gold recovery after fine grinding was 33.9%. All recoveries in this paragraph are as a percent of gold or silver in the concentrate sample tested. The gold recovery after roasting was 72.6%, indicating a significant improvement. In contrast, silver recovery after fine grinding was 82.1%, and silver recovery after roasting was 65.3%, suggesting poorer performance. For comparative Sullivan Gulch samples, gold recovery after fine grinding was 66.5%. Gold recovery after roasting was insignificantly higher at 68.3%. Silver recovery after fine grinding was 92.3%, and silver recovery after roasting was insignificantly lower at 90.1%. The sulfur oxidation in this comparative roaster test was <90%, and that may have contributed to the lack of improvement. The roaster test on the other Sullivan Gulch sample produced gold and silver recoveries of 77.1% and 88.3%, respectively.

Glen Silver Flotation Variability Testing

An additional Glen Silver composite (4307-202) was prepared by combining earlier non-oxide Glen Silver composites 4307-189 thru 4307-194 from drill holes IDM19_117 and IDM19_133 on a footage weighted basis, for grind size optimization flotation tests. Rougher flotation tests were conducted at 80% -150 microns and 80% -75 microns feed sizes. Test results were shown in Appendix B Table 4 (Ref. Test Series 18). Gold and silver recoveries of 80% to 85% reported to flotation rougher concentrates weighing 14.5% to 16.0% of the feed weight. Recoveries were about the same at the two grind sizes. Gold recoveries were moderately higher than obtained from the two other Glen Silver master composites, at the same feed sizes. Silver recoveries fell between those obtained from the two other composites. Sulfide sulfur recoveries were quite high (>97%). The high sulfide recoveries imply that the gold and silver recoveries may be near the maximum achievable by flotation for this Glen Silver composite.

Sullivan Gulch Flotation Variability and Grind-Leach Study

Nine Sullivan Gulch master composites were prepared by combining earlier composites, on a footage weighted basis, according to drill holes, for rougher flotation testing. A rougher flotation test was conducted on each composite at an 80% -75 microns (200 mesh) feed size, using procedures optimized during earlier testing. Summary results from these tests were shown in Appendix B Table 4 (Ref. Test Series 20).

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Gold recovery ranged from 91.0% to 96.8% for seven of the nine composites. Gold recovery for the other composites were 85.8% and 88.2%. Silver recoveries ranged from 96.2% to 99.6% for all but one composite at 85.0%. Mass pulls to rougher concentrate were relatively high and ranged from 9.2% to 27.3% (17.3% average). Sulfide sulfur recoveries were very high (98.5% average) and indicate that the gold and silver recoveries are probably near the maximum achievable by flotation.

A series of cyanidation tests were conducted on “whole ore” feed from 10 of the master composites used for the flotation testing described above, along with another Sullivan Gulch composite (4307-201), to investigate the effects of very fine grinding (“VFG”) and ultra-fine grinding (“UFG”) on gold and silver recoveries by cyanidation. The composites tested were from Glen Silver (one) and Sullivan Gulch (10). The test protocol consisted of a grind-leach cyanidation test on a 1.0 kg sample at an 80% -75 microns (200 mesh) feed size, followed by regrinding and leaching sub-samples split from the grind-leach tailing. Each grind-leach tail sample was used for two regrind-releach tests; one at an approximately 80% -20 microns regrind size (designated very fine grinding) and one at an approximately 80% -11 microns regrind size (designated ultra-fine grinding). Solutions generated during leaching, as well as leached residue from each test and each grind/regrind size were analyzed to determine gold and silver content. Gold and silver recoveries that would be obtained at each of the regrind sizes were calculated by adding the extractions obtained at the 75 microns size to the extraction obtained during the respective regrind-leach test. Particle size distributions (“PSDs”) were determined by laser size analysis, on select leached residues. In the case of the tests where a PSD was not determined, the average of results (P80) from the measured PSDs was assumed (~ 20 microns for VFG and 11 microns for UFG. Results from the test series are shown in Appendix B Table 6 and Figure 13.5.

For the Glen Silver composite, very fine grinding to approximately 80% -20 microns had little or no effect on gold recovery, compared to leaching at the 75 microns feed size. Ultra-fine grinding to approximately 80%-10 microns appeared to improve gold recovery by about 9%, compared to that obtained at 75 microns or 20 microns. Glen Silver composite silver recoveries for VFG and UFG were similar and were significantly higher than the 75 microns grind size.

Silver recovery from the Glen Silver composite improved from 44.3% at the 75 microns feed size to 66.7% for the VFG test. Finer grinding (~ 11 microns) did not further improve silver recovery.

For the Sullivan Gulch composites, gold recoveries for the VFG and UFG tests were similar, and VFG or UFG resulted in a small improvement in gold recovery. Average gold recovery from the Sullivan Gulch composites for the 75 microns (54.0%), VFG (56.5%) and UFG (57.4%) tests were essentially the same. For all three feed sizes, gold recovery tended to increase with increasing CN/FA ratio.

Silver recoveries for Sullivan Gulch composites were significantly improved by very fine grinding to approximately 20 microns in size. At the 75 microns feed size, seven of the 10 samples had silver recoveries between 40% and 50%. The remaining three composites had silver recoveries of 54.3%, 57.6%, and 27.4%. After very fine grinding eight of the 10 composites gave silver recoveries of >78% with an average of 83.7%. The remaining two composites gave silver recoveries of 65.0% and 68.8%. Grinding from approximately 20 microns to 11 microns did not significantly further improve silver recoveries. Average silver recovery for the nine Sullivan Gulch composites was 80% for both the VFG and UFG tests.

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Figure 13.5 Grind-Cyanide Leach Testing Gold and Silver Recoveries, Sullivan Gulch Composites

Flotation Concentrate Albion Processing and Roaster Testing

Glen Silver composite 4307-202 and Sullivan Gulch composite 4307-212 were used to create flotation concentrate for third-party roasting and Albion testing. Five bulk flotation tests, 8.0 kg (17.6 pounds) each, were run to generate sufficient concentrate for these tests (Appendix B Table 4, Test Series 21). The flotation rougher concentrate produced from each test was combined according to composite and blended to create concentrate composites 4307-FC202A and 4307-FC212A. Flotation rougher tailing from each 8.0 kg test were assayed to determine gold, silver and sulfide sulfur grade. Concentrate grades for the flotation tests were determined during the Albion testing.

Flotation test mass pulls were relatively high (14.9% and 19.2%). Gold, silver and sulfide sulfur recoveries were about as expected based on earlier testing on the same composite (Comp. 4307-202) and on earlier testing on the constituent composites comprising the master composite tested (Comp. 4307-212). Silver recovery to the bulk rougher concentrate produced from Glen Silver composite 4307-202 (63.8%) was lower than obtained during an earlier 1.0 kg test (81.7%).

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Roaster tests were performed by Jerritt Canyon Gold (“JCG”) at the Jerritt Canyon mine site laboratory in Elko County, Nevada (Bond, 2020). Baseline CIL tests were conducted on the concentrate after regrinding to 80% -68 microns to 69 microns. Baseline CIL gold recoveries were reported as 20.6% (4307-FC-202A) and 11.7% (4307-212A).

Because of the high sulfide sulfur content of the concentrates (12.6% – 13.4%), the flotation concentrates had to be blended with Jerritt Canyon ore to meet the roaster’s designed operating conditions. Two blends were created for each composite, representing 2% and 5% by weight of the concentrate. Roaster tests performed at JCG resulted in calcine cyanidation gold recoveries exceeding 83%. Silver recovery was not measured. The exact recovery for DeLamar concentrates is not known because the concentrate had to be blended with Jerritt Canyon material to decrease the sulfur level to roaster operating parameters. The blend amounts were 2% and 5% by weight of DeLamar concentrate to respective amounts of Jerritt Canyon ore. In one test, the back-calculated gold recovery for Glen Silver was 93.5%. In the other test, the back-calculated gold recovery was 141.5%, which suggests the added sulfur from the Sullivan Gulch concentrate helped improve recovery from the Jerritt Canyon material.

Samples of the same two concentrate composites were used for scoping level evaluation of the Albion process for recovery of gold and silver from the flotation concentrate. The Albion testing was performed by SGS Minerals in Lakefield, Ontario (Geldart, 2020). The proprietary Albion process employs ultra-fine grinding + atmospheric oxidation followed by agitated cyanidation. Characterization of both concentrate samples included comprehensive assays, specific gravity determination, mineralogy by quantitative XRD analysis, particle size distribution by laser size analysis and a baseline CIL bottle-roll cyanidation test, after ultra-fine regrinding (P80 7.4 to 9.9 microns).

Gold, silver and sulfide sulfur assays from SGS are included in the flotation test results presented in Appendix B Table 4. XRD analysis results indicated that the two concentrates contained primarily quartz (37.8% to 51.3% by weight) and pyrite (16.6% to 19.0%). Lesser amounts of marcasite (3.4% to 7.6%) and arsenopyrite (0.3% to 0.7%) were noted. No other sulfide minerals were detected. The concentrate PSD analyses (laser size analysis – determined on a volume basis) indicated the Glen Silver (Comp. 4307-202) concentrate was 80% -92.5 microns and the Sullivan Gulch (Comp. 4307-212) concentrate was 80% -228 microns. These were significantly different (finer for the Glen Silver composite and coarser for the Sullivan Gulch composite) compared to the flotation test feed size (80% -150 microns or 100 mesh by weight). The reason for that difference was not determined but could include segregation by particle size or attrition during flotation, differences in the methods used (laser size vs. sieving) or sampling error.

The reported Albion testing procedure can be summarized as follows. Aliquots of concentrate, approximately 100 grams (3.53 ounces) each, were reground to the same size as used for the baseline CIL tests (<10 microns). Reground samples were then treated in an agitated glass reactor at a 5% solids density and 95°C while sparging with oxygen. Slurry pH was adjusted to 4.5 with sulfuric acid and the slurry was mixed under these conditions for 60 minutes. Slurry pH was then adjusted to 5.5 using limestone and treatment was continued for a total of 72 hours. Conditions were reported as having been set in consultation with Glencore Technology, owners of the Albion process technology. Summary results for the baseline CIL tests and the Albion oxidation/CIL tests are presented in Appendix B Table 7.

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The Albion process significantly improved gold recovery for both samples. The Glen Silver composite 4307-202 rougher concentrate gold recovery increased from 34.8% to 91.5%. The Sullivan Gulch composite 4307-FC212 rougher concentrate gold recovery increased from 62.5% to 93.0%.

Silver recovery for the Glen Silver composite 4307-202 rougher concentrate increased significantly from 66.6% to 95.0%. Silver recovery for Sullivan Gulch composite 4307-FC212 rougher concentrate increased slightly from 93.4% to 94.2%.

Sulfide oxidation levels achieved during Albion processing were reported as 82.2% and 88.6% for composites 4307-202 and 4307-212 respectively. It is reasonable to presume that the increases in gold recovery result from liberation of gold by sulfide mineral oxidation. It is also likely the case that the improvement in silver recovery from the Glen Silver concentrate resulted also from liberation of silver by sulfide mineral oxidation. The baseline silver recovery by UFG/CIL from the Sullivan Gulch composite was quite high (93.4%). Silver recovery from this concentrate did not benefit significantly from the Albion process.

CIL reagent consumptions for the Albion tests were reported as 1.25 to 1.57 kg/t NaCN and 4.98 to 5.73 kg/t CaO. These were reported on a flotation concentrate weight basis.

Considering data from the flotation tests used to generate these concentrates, the Albion recoveries obtained from the Glen Silver composite were equivalent to 73% gold and 76% silver, on a flotation feed basis. The Albion recoveries obtained from the Sullivan Gulch composite were equivalent to 84% gold and 86% silver on the same basis. These results demonstrate that high gold and silver recoveries can be achieved from the DeLamar flotation concentrates by Albion processing.

Flotation Concentrate Regrind/Cyanidation Testing

A total of 12 DeLamar non-oxide drill core composites were prepared for flotation testing and flotation regrind/cyanidation testing to evaluate response to processing optimized for the DeLamar non-oxide material. Ten of the twelve composites were comprised of material from the Sullivan Gulch area. Two of the Sullivan Gulch composites (4522-075 and 076) represented the Tsgl lithology. A third composite (4522-074) was comprised of multiple lithologies with over half being Tsgl material. The remaining seven composites represented various other lithologies from the Sullivan Gulch area. The Sullivan Gulch composites were selected from material having relatively low CN/FA ratios averaging 8.6%. The two remaining composites represented material from the Glen Silver area (4307-193/194) and from the Sommercamp area (4522-085).

Grind-leach tests were conducted on 11 of the 12 composites at an 80% -75 microns (200 mesh) feed size to confirm their refractory nature. Average gold and silver recoveries from the SG composites were 17.7% and 47.7%, respectively. Composite 4307-193/194 was not tested.

Two flotation tests were conducted on each composite, using conditions optimized during earlier testing, to confirm response to flotation treatment and to generate concentrate for concentrate regrind/-cyanidation tests. Feed size for the flotation tests was 80% -150 microns (100 mesh). Only one test was possible on composite 4307-193/194, because of sample limitations. Flotation test results were presented in Appendix B Table 4 (Test Series 4522). Results showed that the non-Tsgl lithology Sullivan Gulch composites responded reasonably well to sulfide flotation treatment. Gold values reporting to the flotation rougher concentrate (11.0% average mass pull) ranged from 67.2% to 94.5% and averaged 85.1% of the total gold. Silver reporting to the same rougher concentrates ranged from 74.2% to 94.5% and averaged 87.8%. Sulfide sulfur recoveries to the rougher concentrates generally were greater than 90% and averaged 94.9%.

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The three composites containing predominantly Tsgl material (4522-074, 075, 076) responded more poorly to flotation treatment. Mass pulls and recoveries were variable. Problems related to gangue carry-over to the rougher concentrate were observed. Gold recoveries to the Tsgl composite rougher concentrates ranged from 40.6% to 81.6% averaging 57.4%. Silver recoveries ranged from 62.8% to 85.4% averaging 74.8%. Sulfide sulfur recoveries ranged from 77.4% to 95.8% averaging 86.4%. Results indicate that further optimization of flotation conditions will be required to maximize recoveries from this material type. It was noted by Integra resources personnel that the Tsgl lithology represents a relatively small portion (3% as indicated by drill hole footage) of “ore grade” non-oxide material from the DeLamar deposit.

The Glen Silver composite (4307-193/194) responded relatively poorly to flotation treatment under the conditions evaluated. Only a single test was possible. The flotation rougher concentrate produced was 16.8% of the feed weight and represented gold and silver recoveries of 47.3% and 53.8% respectively. Results from this flotation test are not believed to be as representative of the Glen Silver non-oxide material as the master composites tested during earlier work (Wickens, 2020).

The Sommercamp composite (4522-085) responded well to flotation treatment. The flotation test used to generate the concentrate for regrind/cyanidation had an anomalously low mass pull (6.5%). The other flotation concentrate produced was 10.4% of the feed weight and represented gold and silver recoveries of 89.1% and 93.1%, respectively. Sulfide sulfur recovery for that test was 92.6%.

Agitated cyanidation tests were conducted on flotation concentrate generated from each of the composites after fine regrinding. Regrind sizes ranged from 80% -7.0 microns to 80% -39 microns and averaged 80% -17.7 microns. Those results are shown in Table 13.4.

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Table 13.4 Flotation Concentrate Regrind/Cyanidation Tests, DeLamar Non-Oxide Composites

The concentrates were not particularly amenable to cyanidation for recovery of contained gold. Gold recovery obtained from the Sullivan Gulch concentrates ranged from 11.1% to 38.0% and averaged 23.9%. These recoveries were equivalent to between 4.5% and 30.7% and averaged 18.5% of gold contained in the flotation feed. This compares to an average CN/FA for these composites of 8.8% gold.

Gold recoveries obtained from the Sommercamp and Glen Silver concentrates were somewhat higher at 44.5% and 61.6%, respectively, and were equivalent to 33.2% and 29.1%, respectively, of gold contained in the flotation feed. Respective CN/FA ratios for these composites were 19.2% and 23.9%. Concentrate cyanidation gold recoveries were not sensitive to regrind size, within the range evaluated.

Silver recoveries obtained from the Sullivan Gulch concentrates were consistently higher and ranged from 67.9% to 88.0% with an average of 82.6%. These recoveries were equivalent to between 46.6% and 67.3% (56.2% average) of silver contained in the Tsgl composites and to between 62.0% and 77.8% averaging 72.5% of silver contained in the other Sullivan Gulch composites. Silver recovery from the Sommercamp concentrate was quite high at 94.2% and was equivalent to 84.8% of silver contained in the flotation feed. Silver recovery from the Glen Silver composite was lower at 53.3% and was equivalent to 28.7% of silver contained in the flotation feed.

Reagent consumptions for the concentrate leach tests generally were moderate. Cyanide consumption for the Sullivan Gulch concentrates ranged from 2.94 to 22.53 kg NaCN/t concentrate, which was equivalent to between 0.33 and 2.30 kg NaCN/t of flotation (“whole ore”) feed. Lime required for the Sullivan Gulch concentrates was equivalent to between 3.1 and 25.0 kg/t concentrate, which was equivalent to between 0.3 and 2.6 kg/t flotation feed. Cyanide leaching conditions were not optimized, and it is expected that reagent consumptions can be decreased through optimization testing.

Scoping level flotation tests with regrind-agitated cyanidation tests on the rougher concentrates were conducted on the four DeLamar mixed samples used for column-leach testing. The tests were conducted to assess the potential for processing the mixed material by milling. Flotation recoveries from the mixed composites generally were poor and averaged 51.6% gold and 62.7% silver. Further testing was recommended to evaluate improving flotation response for the mixed material. Concentrate cyanidation gold recoveries varied from 41.3% to 91.7% of gold contained in the concentrate. Corresponding silver recoveries averaged 93.1% of silver in the concentrate. Regrind sizes were finer than targeted and ranged from 80% -9.82 microns to 80% -5.05 microns. Additional testing would be required to determine recoveries at the planned regrind size of 80%-20 microns but were not recommended unless flotation response was first improved.

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13.3.1.6DeLamar Comminution Testing

A total of six comminution composites were prepared from 2020 DeLamar drill core during the PFS testing for crusher work index, abrasion index and ball mill work index testing. Crusher work index tests were run using a Bond low energy impact test via twin pendulums. Abrasion index tests were conducted using a Bond impellor-tumbler impact abrasion test method. Ball mill index testing was conducted using standard Bond methods. Drill core samples were selected to represent the major lithology types from the deposit. Results from the comminution tests are shown in Appendix B Table 8.

Crusher work index testing classified all material types tested as very soft. Bond crusher work indices ranged from 3.92 to 8.26 kWh/t (5.88 kWh/t average). Bond abrasion indices were variable and classified as lightly abrasive to very abrasive. The Glen Silver composites were classified as very abrasive. Bond ball mill work indices were variable and classified as medium to very hard. Ball mill work indices were highest for the Glen Silver composites (23.98 to 24.86 kWhr/t) and ranged from 11.36 to 19.45 kWhr/t for the other composites.

13.3.2Integra Florida Mountain Area Testing

Testing at McClelland on the oxide and mixed material types included column-leach testing at multiple feed sizes on a total of 18 drill core composites. Detailed head analyses and a comparative bottle-roll test (80% -1.7mm feed size) (10 mesh) were conducted on each sample. In addition, a total of 67 variability core composites were prepared for the same bottle-roll testing and head analyses. Load/permeability testing was conducted on select column residues to evaluate permeability under simulated commercial heap stack heights.

Testing on non-oxide material included evaluation of gravity concentration, whole ore grind/leach and flotation with flotation concentrate processing by regrind/cyanidation. Gravity concentration testing and most of the flotation optimization work was done during the PEA testing. Flotation concentrate processing testwork was done mostly during the PFS testing.

13.3.2.1Florida Mountain Samples

Composites were prepared from a total of 30 drill holes, eight of which were from 2018 drill holes and 22 of which were from 2019 drill holes. The composites from the 2018 drilling were used for the PEA testing. The PFS testing was conducted on composites from the 2019 drilling. Composites tested included 24 oxide, 56 mixed, 31 non-oxide and three blended (multiple oxidation classes) material type composites. A summary of the drill core samples used for compositing is shown in Table 13.5. A map of the drill hole collar locations for the metallurgical drill holes are shown in Appendix B Figure 2.

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Detailed head analyses conducted on each of the metallurgical composites included fire assay to determine gold and silver content, cyanide shake analysis to determine cyanide soluble gold and silver, carbon and sulfur speciation analyses and a multi-element ICP scan. Head assays conducted on the composites from the 2019 drilling were conducted using conventional “metallic screen” fire assay procedures. This procedure was used because of the relatively high level of head grade variability observed with the Florida Mountain material. The metallic screen assaying entailed pulverization of a 1.0 kg sample (>95% passing 106 microns) (150 mesh), screening of the pulverized sample on a 106 microns sieve, fire assaying the oversized material (>106 microns) to extinction and assaying duplicate spits of the undersized (<106 microns) material. Assaying in this manner is an effective means for mitigating excessive grade variability and allows for evaluation of the presence of coarse particulate gold by consideration of the gold distribution in the coarser and finer fractions.

Head assays showed the composites contained between 0.05 and 6.71 g Au/t (0.69 g Au/t average) and between 2.0 and 343 g Ag/t (28 g Ag/t average). Average gold head grades for the oxide, mixed and non-oxide composites were 0.77, 0.71 and 0.60 g Au/t, respectively. Respective average silver head grades were 32, 33 and 28 g Ag/t.

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Table 13.5 Drill Hole Composite Summary, Florida Mountain PEA and PFS Testing

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Head grade agreement generally was acceptable to poor for the Florida Mountain composites. In 10 of the 57 composites tested by metallic screen head assay, 30% or more of the total contained gold reported to the metallics fraction. This fraction averages less than 1% of the sample weight. This degree of upgrading in the metallics fraction is indicative of the presence of “coarse particulate gold” and is expected to cause difficulties with head grade variability. Head grade relative standard deviation for the PFS column test composites ranged from 9% of the average head grade to 75% of the average head grade (25% average). Head grade agreement problems were mitigated to a significant degree by using metallic screen fire assaying procedures, 1.0 kg samples, rather than conventional 30-gram (1.06-ounce) fire assay procedures, on products from the metallurgical tests, where possible and appropriate. Similar, but less severe, variability in silver head grades was also observed for the Florida Mountain material. Head grade variability can be expected to be a significant challenge for the Florida Mountain deposit.

CN/FA soluble to fire assay gold ratios for the oxide, mixed and non-oxide composites averaged 58.9%, 64.4% and 65.5%, respectively. Cyanide shake analyses were poorly correlated to cyanide leach amenability, as measured during metallurgical testing. This is believed to be in part because of the high degree of head grade variability observed with the Florida Mountain material. Cyanide shake analysis will likely not be a useful analytical method for ore control for the Florida Mountain deposit.

Average sulfide sulfur head grades for the oxide, mixed and non-oxide composites were 0.03%, 0.22% and 1.01%, respectively. Sulfide head grades for the oxide and non-oxide material types were reasonably consistent when comparing the two sets of drill hole samples (2018 versus 2019). Sulfide sulfur grades for the mixed material type were on average significantly higher for the 2019 drill hole composites (0.31% S) compared to the 2018 drill hole composites (0.06% S average). As heap-leach amenability (gold recovery by bottle-roll or column testing) for the mixed material composites was correlated to sulfide sulfur grade, the difference in grades for the two drill campaigns was reflected in generally lower gold recoveries from the mixed material for the 2019 drill hole composites, compared to the 2018 drill hole composites. Average sulfate sulfur grades were 0.04% S (oxide), 0.05% S (mixed) and 0.22% S (non-oxide). Organic (non-carbonate) carbon content generally was negligible (<0.1%) and no indications of preg-robbing were observed during any of the metallurgical testing.

The Florida Mountain composites contained low to negligible concentrations of potentially deleterious elements, other than sulfur. Only two of the 106 composites tested contained greater than 100 ppm Cu. Average copper grades of the oxide, mixed and non-oxide composites were 9, 14 and 38 ppm, respectively. Respective average mercury head grades were 0.06, 0.05 and 0.05 ppm. Selenium concentration did not exceed 20 ppm and averaged 3.0 to 4.0 ppm for the three oxidation classes.

13.3.2.2Florida Mountain Mineralogy

As mentioned in Section 13.3.1.2, a total of 67 metallurgical samples from the McClelland PEA and PFS testing program were submitted to Vidence, Inc., in Burnaby, British Columbia for automated SEM scans to determine the mineralogy and texture of the materials (Enter, 2021). The samples included 20 from the Florida Mountain deposit, of which three were oxide, 10 were mixed and seven were non-oxide material type samples. The remaining samples were from the DeLamar area deposit discussed in Section 13.3.1.2.

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Florida Mountain samples were found to have moderately variable quartz and feldspar content with quartz ranging from 42% to 73% and feldspar ranging from 9.5% to 47%. Plagioclase was sporadically distributed varying from absent to 10%. Sulfide minerals, as expected, were most common in the non-oxidized mineralization with trace amounts present in the mixed mineralization. Pyrite was the most common sulfide identified, up to 2.2%, with trace amounts of chalcopyrite, up to 0.11%. Jarosite was the only sulfate identified and was most abundant in the mixed mineralization, up to 0.29%. The oxide mineralization was found to contain trace to no sulfides and only trace amounts of jarosite.

Clay species were found to be moderately diverse and locally abundant. Muscovite and illite were found to be the main clay species ranging up to 14% and 9.1%, respectively. Lesser amounts of kaolinite, up to 4.5%, and Al-silicate clays, up to 0.5%, were also identified. Chlorite was found sporadically, specifically in the non-oxidized mineralization, at abundances up to 13%. A strong correlation exists between lithology and clay mineral assemblage. Muscovite is most abundant in samples where plagioclase is absent or small amounts with moderate amounts of K-feldspar. Observed kaolinite correlated with the abundance of plagioclase. Both of these correlations were spread across all three oxidation classifications.

13.3.2.3Integra Florida Mountain Heap-Leach Testing

A detailed heap-leach testing program was conducted in two phases on oxide and mixed samples from the Florida Mountain deposit. Phases consisted of the PEA work conducted on 2018 drill samples and the PFS work conducted on 2019 drill samples. For the ease of discussion, results from both programs are presented together here where appropriate.

Both phases of testing included bottle-roll tests on drill core variability composites (67 total). The PEA testing composites (2018 drilling) were comprised of 28- to 108-meter (92- to 354-foot) intervals. The PFS composites generally were comprised of 4.6 to 7.3 meters (15 to 24 feet) of continuous interval. Three of the PFS composites were comprised of longer intervals of core (25 to 53 meters) (82 to 174 feet). The variability composites included 28 oxide and 38 mixed type composites. Column leach tests were conducted on a total of 18 drill core composites, including four oxide, 13 mixed and one non-oxide. Tests were conducted at an 80% -12.7mm on 15 composites. Comparative tests were conducted on eight of the same composites at a -50mm feed size. Two additional core composites were tested only at a -50mm feed size. A single mixed material type composite was column tested after preparation to a finer feed size, using high HPGR. Tests were conducted in a manner to determine gold and silver recoveries, recovery rates and reagent requirements under simulated heap leaching conditions. A comparative bottle-roll test was conducted on each of the column test samples at the same size as used for the variability tests (80% -1.7mm) (-0.5-inch) to help establish the relationship between bottle-roll and column test recoveries.

Heap-leach Variability Bottle-Roll Testing

Bottle roll tests were conducted on each of the variability composites, at an 80% -1.7mm feed size, to evaluate potential for heap leaching and material variability. An additional 20 non-oxide composites were also tested using the same procedures. The tests were conducted using 4-day leach cycles with a cyanide concentration of 1.0 g NaCN/L. Summary results from bottle-roll tests on the oxide and mixed variability composites are presented in Appendix B Table 9 and Appendix B Table 10, and are presented graphically in Figure 13.6.

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Figure 13.6 Florida Mountain Bottle-Roll Test Recoveries, Variability Composites

The oxide composites were readily amenable to agitated cyanidation treatment at the -1.7mm (-0.5-inch) feed size. Gold recoveries from the 28 composites ranged from 68.7% to 97.2% and averaged 85.5% in four days of leaching. Only five of the 28 composites tested gave gold recoveries of <80%.

Silver recoveries obtained from the oxide composites were highly variable and generally lower compared to the gold recoveries. Silver recovery from the 26 composites ranged from 14.3% to 87.5% and averaged 47.2% in four days of leaching. One-half of the composites gave silver recoveries of <45%.

A total of 38 mixed composites were tested. Overall, gold recoveries obtained in four days of leaching ranged from 40.6% to 97.0% and averaged 75.8%. About one-half (20 of 38) of the mixed composite gold recoveries were lower than 80%. Although gold recoveries from the mixed material were not strongly correlated to composite sulfide grade, the PEA (2018 drilling) mixed composites tended to have lower sulfide grades (0.08% average) and higher gold recoveries (81.4% average) compared to the PFS (2019 drilling) mixed composites (0.35% sulfide sulfur and 73.6% Au, average).

Silver recoveries from the mixed material were variable and ranged from 25.0% to 84.2% (48.6% average). Silver recoveries tended to be higher on average (52%) for the higher sulfide grade PFS composites compared to the lower sulfide grade PEA composites (40.0% average). Silver recovery was not strongly correlated with sulfide sulfur grade.

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Recoveries obtained from the non-oxide composites generally were low and averaged 51.4% gold and 38.0% silver.

Reagent consumptions were low. Cyanide consumptions for the oxide and mixed composites averaged 0.13 and 0.22 kg NaCN/t, respectively. None of the oxide composites and only three of the mixed composites gave cyanide consumptions greater than 0.3 kg NaCN/t. Average lime requirement was 0.9 kg/t for both the oxide and mixed material types.

Florida Mountain Column-Leach Tests

Seven column tests were conducted in support of the 2019 PEA using composites of 2018 drill core samples. A more extensive column testing program (19 column tests total) was conducted in support of the PFS using composites of 2019 drill core samples. All columns conducted for the PEA were run at an 80% -12.7mm (-0.5-inch) feed size. Columns conducted for the PFS study were run at a -50mm (-2-inch) feed size (10 tests), and at an 80% -12.7mm feed size (eight tests). A single test was also run on a mixed master composite of 2019 drill core samples, after single-pass HPGR crushing to a finer (77% -6.3 mm) (-0.25-inch) feed size.

All column tests were conducted without agglomeration pretreatment, but with hydrated lime (0.7 to 3.4 kg/t) blended with the feeds before leaching for pH control. Leaching conditions included solution application at a rate of 9.8 Lph/m2 at a cyanide concentration of 1.0 g NaCN/L for leach cycles ranging from 63 to 97 days for the PEA testing and from 81 to 240 days for the PFS testing. Tests were conducted to determine gold and silver recovery, recovery rate and reagent consumptions under simulated heap-leaching conditions.

Comparative bottle-roll leach tests were conducted on each composite that was column tested, at an 80% -1.7mm (10 mesh) feed size, using essentially the same conditions as described previously for the variability sample testing. The purpose of these tests was to establish the relationship between bottle-roll and column test recoveries. Kinetic, recovery versus time, sampling and analysis procedures were performed on the column bottle-roll tests, but generally not on the variability bottle-roll tests.

Column test results from both testing programs are summarized in Table 13.6, Appendix B Figure 3 and Appendix B Figure 4.

The Florida Mountain material was variable in response to simulated heap-leaching treatment (column leaching). The oxide composites generally gave very high gold recoveries (89.6% average at -12.7mm). Gold recoveries from the mixed composites were variable, with the low sulfide (<0.1% S) mixed composites giving very high gold recoveries (similar to the oxides) and the elevated sulfide (0.1% - 0.5% S) mixed composites giving gold recoveries that ranged from about 60% to 80%. It was also noted that the low sulfide mixed composites in the PEA testing tended to be from shallower material and were all "Tpr" lithology composites. Most of the elevated sulfide mixed composites of the PFS testing tended to be from deeper material and were mostly "Tql" lithology composites. As expected, the single non-oxide composite that was column tested gave a relatively low gold recovery (30.0%).

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Table 13.6 Column-Leach and Bottle-Roll Test Results, Florida Mountain PEA and PFS Testing

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Column test silver recoveries from the -12.7 mm (-0.5-inch) feeds were significantly lower and not as variable. Silver recoveries from the oxide composites ranged from 41.3% to 52.9% and averaged 47.6%. Silver recoveries from the mixed composites ranged from 26.3% to 54.5% and averaged 41.8%. Silver recovery from the non-oxide column test was only 10.0%.

As shown in Appendix B Figure 4, the oxide composites all contained relatively low sulfide sulfur levels (<0.1%) and gave consistently high gold recoveries. The mixed and blended (mix/ox) composites from the PEA testing program all contained relatively low levels of sulfide sulfur (<0.1%) and gave relatively high gold recoveries (>85%). The mixed composites leached at the -12.7mm (-0.5-inch) feed size during the PFS testing and one of the blended (mix/non-oxide) composites from the PEA program contained higher levels of sulfide sulfur (0.14% – 0.39% S) and gave lower gold recoveries (62.5% – 76.0%). It is believed that some of the low sulfide material logged as mixed during 2018 drilling may be better classified as oxide material, and that the mixed material tested during the PFS study better represents the mixed material. This information shows a need for further refinement of the oxidation class logging and modelling.

Neither gold nor silver recovery from the Florida Mountain samples were strongly correlated to lithology, head grade or any of the trace metal analyses conducted on each of the composites. As noted above, there was a tendency for the Tpr lithology mixed composites to contain relatively low sulfide sulfur levels and to give relatively high gold recoveries. It is believed to be more likely that the recovery variability is related to sulfide grade than lithology. This observation is supported by the similar average variability bottle-roll test gold recoveries obtained from the two lithology types.

Gold and silver leach rates for the mixed composites also tended to be slow , but in some cases were more rapid. In most cases, gold and silver extraction would be expected to improve incrementally with longer leach times.

Select composites (eight total) were tested at both a -50 mm (-2-inch) feed size and at the 80% -12.7mm feed size to optimize heap-leach feed size. In the case of the oxide composites, the -50mm feed size tests were stopped before leaching was completed, after about 70 to 80 days of leaching. The corresponding -12.7mm tests were run to completion. Gold recovery rate tended to be slower at the coarser feed size and comparative rate data indicate that, allowed sufficient leaching time, gold recovery achieved at both feed sizes would be very similar. In all three cases (oxides) it was not clear that the approximately 10% difference in silver recovery between the two feed sizes would narrow significantly with more leaching time. This indicates potential for leaching of the Florida Mountain oxide material at a coarser feed size while maintaining gold recovery levels achieved at -12.7mm, but likely with moderately lower silver recoveries.

In the case of the mixed material composites tested at both feed sizes, the results were less consistent. In some cases (e.g., composite 4471-052), a large decrease in gold and silver recovery would be expected from coarsening the crush size from -12.7mm to -50mm. In other cases, (e.g., composite 4471-053), a similar ultimate gold recovery would be expected at the two feed sizes, with silver recovery expected to be substantially lower. These results indicate that the Florida Mountain mixed material is moderately to strongly sensitive to feed size for heap-leach gold and silver recoveries and requires three stages of crushing for optimal heap-leach recoveries.

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A single master composite (4471-069) was prepared from the higher-grade mixed column test composites (4471-053, 4471-055 and 4471-057) on a footage weighted basis. This composite was used for a column test at a finer feed size (77% -6.3mm) (-0.25-inch) after crushing by single-pass through pilot HPGR. The HPGR sample preparation was conducted at Kappes Cassidy and Associates (“KCA”), in Sparks, Nevada (KCA, 2021). Column test results for the HPGR product showed a gold recovery (73.3%) which was approximately the same as would be expected at the -12.7mm (-0.5-inch) feed size, based on column test results from the samples comprising this feed. Silver recovery obtained from the HPGR product (63.6%) was about 12% higher than expected at a -12.7mm feed size. Replicate testing would be required to confirm the indicated improvement in silver recovery, in part because of the low-grade nature of the composite (11 g Ag/t).

Head and tail screen analyses were conducted on the column feeds and residues to help evaluate feed size sensitivity. Residues from the oxide tests at -12.7mm were quite low in gold grade (0.03 to 0.04 g Au/t). Residues from the mixed composite tests at -12.7mm were more variable in gold grade (0.06 to 0.29 g Au/t). Recovery by size data were made difficult to evaluate because of the degree of assay variability encountered with the column feeds and residues. Extensive use of metallic screen assaying was made on screen size fractions to help ensure accurate grade determinations. Recovery by size data tended to be erratic but generally indicated fine grinding would be required to maximize recoveries. Silver assays also tended to be erratic, but not to the same degree as the gold assays.

Cyanide consumptions for the PEA composites leached for 63 to 65 days ranged from 1.16 to 1.29 kg NaCN/t. Cyanide consumptions for the two composites leached for 97 days ranged from 2.03 to 3.08 kg NaCN/t. Cyanide consumption for the much longer duration PFS column tests were substantially higher and ranged from 1.59 to 2.93 kg NaCN/t for the oxide composites and from 1.71 to 2.98 kg NaCN/t for the mixed composites. Considering the low bottle-roll test cyanide consumptions (0.16 kg NaCN/t average) and column test cyanide consumptions through a traditional 60-day column-leach cycle (1.2 to 1.3 kg NaCN/t average), it is anticipated that commercial cyanide consumptions for the oxide and mixed material types would not exceed 0.4 and 0.6 kg NaCN/t ore, respectively. The 0.5 to 3.4 kg/t lime, added before leaching, was sufficient for maintaining leaching pH. Lime consumptions equivalent to 2.0 kg CaO/t (oxides) and 2.4 kg/t (mixed) were estimated for commercial heap leaching.

Column test gold recoveries (-12.7mm feed size) and bottle-roll test gold recoveries (-1.7mm feed size) (10 mesh) for the Florida Mountain material were strongly correlated, demonstrating that short term bottle-roll testing on fine sample (i.e., crushed drill core, RC cuttings or blast-hole cuttings) will be useful for predicting heap-leach gold recoveries. Silver recovery data (column- versus bottle-roll tests) were not as strongly correlated and would be of limited use for predicting heap-leach silver recoveries.

All the Florida Mountain samples subjected to column leaching displayed good solution percolation and permeability characteristics during column leaching. The degree of ore slump observed during leaching was minimal.

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Florida Mountain Load-Permeability Testing

Select column-leached residues (-12.7mm feed size) generated during the PEA and PFS testing were submitted to Geo-Logic Associates, in Sparks, Nevada, for load/permeability tests (test procedure USBR 5600-89). The tests were conducted to evaluate permeability of the leached material under expected commercial heap stack heights. Results showed generally high permeabilities over a range of simulated heap stack heights of up to 137 meters (449.5 feet) (McPartland, 2022). Only one of the three oxide composites had a permeability of less than the equivalent solution application rate of 6.0 Lph/m2 (1.65E-04 cm/s) and that was only at stack heights of 122 meters (400 feet) or greater (final permeability was 1.02E-04 cm/s at 137 meters) (449.5 feet). Otherwise, for both the oxide and mixed -12.7mm (-0.5-inch) feeds, permeability was equivalent to approximately 10 times or greater than the anticipated solution application rate. These results demonstrate that the Florida Mountain oxide and mixed materials generally will not require agglomeration pretreatment for successful heap leaching at heap stack heights of up to about 150 meters (492 feet).

13.3.2.4Integra Florida Mountain Mill PEA Testing

Testing in support of the 2019 PEA was conducted mainly on 10 non-oxide drill core composites and one overall “master” non-oxide drill core composite. That testing included “whole ore” grind/agitated cyanidation (bottle-roll) tests, gravity concentration followed by cyanide leaching of the gravity tails, and gravity concentration followed by flotation and regrind/cyanidation of the flotation concentrate.

Florida Mountain Grind-Leach Tests

Ten of the 2018-2019 Florida Mountain non-oxide composites, along with one of the mixed composites were used for bottle-roll leach tests at an 80% -75 microns (200 mesh) feed size to evaluate amenability to grind-leach (cyanidation) treatment. These tests were conducted using a 72-hour leach cycle, at a solids density of 40% and a cyanide concentration of 1.0 g NaCN/L.

All of the non-oxide Florida Mountain composites tested were amenable to grind-leach treatment, at an 80% -75 microns feed size. Gold recoveries ranged from 76.8% to 96.1% and averaged 85.7% in 72 hours of leaching. Corresponding silver recoveries ranged from 32.7% to 90.8% and averaged 61.5%. Reagent consumptions were fairly low.

The single Florida Mountain mixed composite tested gave very high recoveries of 94.4% gold and 92.5% silver. Further testing of Florida Mountain mixed material for grind-leach and gravity concentration /tailing cyanidation processing will be required for a trade-off study against heap leaching.

Florida Mountain Gravity Concentration, Treatment of Gravity Tailing

Considering the consistent behavior of the Florida Mountain non-oxide composites, a single master composite was prepared for comparative gravity concentration with evaluation of gravity-tailing cyanidation and gravity-tailing flotation. The resulting flotation concentrate was also subjected to regrinding followed by intensive cyanidation testing. The composite (designated 4307-160) was a “master composite” prepared from non-oxide composites 4307-135, 140, 141 and 142, which were comprised of drill core from holes IMF18_003, 004 and 010.

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A single gravity-concentration test was conducted on the Florida Mountain non-oxide master composite (4307-160) to evaluate response to gravity concentration and to generate gravity tailing for cyanidation and flotation testing. The resulting cleaner concentrate was assayed. The cleaner tails and rougher tails were recombined and then split to obtain feeds for cyanidation and flotation testing. The gravity cleaner was 0.04% of the feed weight, assayed 148 g Au/t and 316 g Ag/t, and was not included in the agitated cyanidation or flotation test feeds.

Agitated cyanidation tests were conducted on gravity tailing, generated as described in the preceding paragraph, at five tailing regrind sizes ranging from 80% -150 microns (100 mesh) to 80% -45 microns (325 mesh). The gravity tailing was amenable to agitated cyanidation treatment at the regrind sizes tested. Combined gold recovery (gravity concentration + tailing cyanidation) ranged from 80.9% and 82.2% and was not sensitive to regrind size. Combined silver recoveries increased with decreasing regrind size, from 57.0% to 71.3%. Reagent consumptions were low.

Bulk sulfide flotation tests were also conducted on separate splits of the same gravity tailing, at regrind sizes ranging from 80% -212 microns (65 mesh) (no regrind) to 80% -75 microns (200 mesh), to evaluate the potential for upgrading the gravity tailing by flotation. Summary results from the gravity/flotation tests are shown in Table 13.7.

Table 13.7 Florida Mountain 2018-2019 Gravity Concentration with Flotation of Gravity Tailing

(Comp. 4307-160, 80%-212 microns Gravity Concentration Feed)

The gravity tailing also responded very well to bulk sulfide flotation. The combined concentrates (gravity cleaner concentrate + flotation rougher concentrate), produced at regrind sizes of as fine as 106 microns, were equivalent to between 6.6% and 11.2% of the “whole-ore” mass, and contained 94.9% to 97.6% of the gold and between 87.5% and 89.8% of the silver contained in the “whole-ore”. Although mass pull to the flotation concentrate tended to increase with decreasing regrind size, recoveries did not increase. Sulfide sulfur recovery by flotation ranged from 86.4% to 90.4%. Gold recovery (90.1%) and sulfide sulfur recovery (76.9%) were somewhat lower at the 75 microns regrind size.

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Florida Mountain Flotation Concentrate Regrind/Agitated Leach

Based on the positive results obtained from the flotation testing conducted on the Florida Mountain non-oxide material, a larger, 8-kilogram flotation test was conducted on the same gravity tailing (non-oxide master composite 4307-160), at an 80% -212 microns feed size (no regrind) to generate rougher concentrate for cyanidation testing. The purpose for this test was to determine if, by fine regrinding of the flotation concentrate, cyanidation recoveries could be improved beyond those observed during agitated cyanidation testing on the same gravity tailing. The flotation rougher concentrate produced was used as feed for an intensive cyanidation test. Summary test conditions (flotation concentrate leach) and summary gravity/flotation/concentrate cyanidation test results are summarized in Table 13.8

Table 13.8 Florida Mountain 2018-2019 Gravity, Flotation of Tailing and Regrind Leach of Flotation Concentrate

(Unoxidized Master Composite 4307-160, 80% -212 microns Feed Size; 95% -37 microns Flotation Concentrate Regrind, 96 Hour Leach, 25% Solids, 5.0 g NaCN/L)

The flotation rougher concentrate produced from the Florida Mountain non-oxide master composite 4307-160 was readily amenable to agitated cyanidation treatment, at a 95% -37 microns regrind size. Gold and silver recoveries were 93.1% and 89.8%, respectively, from the flotation concentrate. The combined recoveries by gravity concentration (cleaner concentrate) and cyanidation of the flotation concentrate were equivalent to 89.7% of the gold and 80.2% of the silver contained in the “whole-ore” feed. Cyanidation recovery rates were moderate, and extraction of gold and silver was progressing at a slow, but significant rate when leaching was terminated after 96 hours. Optimization of the leaching cycle and leaching conditions will likely lead to incrementally higher recoveries. Cyanidation reagent consumptions were very low and equivalent to only 0.14 kg NaCN/t and 0.2 kg lime/t on a “whole-ore” mass basis.

Recoveries obtained by regrind and cyanidation of the flotation concentrate compare favorably to those obtained by gravity concentration with cyanidation of the gravity tailing. Gold and silver recoveries obtained from the master composite by gravity concentration with cyanidation of the gravity tailing did not exceed 82.2% and 71.3%, respectively, at gravity tailing regrind sizes as fine as 80% -45 microns (325 mesh). These results indicate apparent increases in overall gold and silver recoveries of approximately 7% and 9%, respectively, were obtained by very fine regrinding of the flotation concentrate. The test results demonstrate that a relatively coarse primary grind size, with correspondingly lower grinding costs, will be possible for the Florida Mountain non-oxide material.

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13.3.2.5Florida Mountain Mill PFS Testing

Detailed mill testing was conducted on a total of 10 non-oxide composites from the Florida Mountain deposit, to optimize processing conditions for recovery of gold and silver by flotation and agitated cyanidation of the flotation concentrate, and to evaluate ore variability. Initially seven composites were prepared from five drill holes (205.7 lineal meters of drill core total) (675 lineal feet), including two from the Tip-Top area and three from the North to Central portion of the Main area and one from the South to Central portion of the Main area.

Detailed head analyses conducted on the composites included metallic screen assay to determine gold and silver content, cyanide shake analysis to determine cyanide soluble gold and silver, carbon and sulfur speciation analyses and a multi-element ICP scan. Head assays showed the composites contained between 0.14 and 0.42 g Au/t (0.32 g Au/t average) and between 3 and 33 g Ag/t (8 g Ag/t average). Two of the seven composites were described potentially containing relatively coarse particulate gold values, with 39% to 59% of the contained gold reporting to a relatively high grade “metallics fraction” during metallic screen assaying. Cyanide soluble to fire assay (“CN/FA”) gold values were erratic and shown with more detailed testing to not be a reliable indicator of cyanide leach amenability. The composites contained 0.54% to 2.28% sulfide sulfur (1.21% average). Sulfate sulfur levels were much lower (0.21% average).

Agitated Leach (Bottle-Roll) Tests

Scoping agitated cyanidation tests were conducted on each composite at 80% -1.7mm feed size and at an 80% -75 microns (200 mesh) feed size. Each were leached for 96 hours, using a solids density of 40% and a cyanide concentration of 1.0 g NaCN/L.

Bottle roll test results showed that the composites were amenable to agitated cyanidation treatment. Gold recoveries (44.4% – 70.9%) and silver recoveries (30.2% – 50.0%) obtained at the 1.7mm (10 mesh) feed size indicate some potential for heap leaching of the non-oxide material but expected recoveries would be significantly lower than obtained from oxide or mixed materials. Comparative bottle-roll test recovery results for samples from both the PEA and PFS studies for the three oxidation classes were presented graphically in the heap-leach testing section, in Figure 13.6. (Sect. 13.3.2.3).

Results from the grind-leach cyanidation tests confirmed that the non-oxide material was not refractory to cyanidation treatment. Gold recoveries obtained in 4 days of leaching ranged from 70.0% to 89.5% (81.9% average). Silver recoveries ranged from 43.8% to 63.0% (55.4% average).

Gold and silver recovery rates for the 1.7mm feeds were moderate to fairly slow, and in most cases extraction was progressing at a slow rate when leaching was terminated. Recovery rates were more rapid for the 75 microns feeds and gold leaching generally was complete in 96 hours or less. Silver extraction generally was continuing at a slow rate when leaching was stopped.

Cyanide consumptions were fairly low and averaged 0.3 kg NaCN/t at both feed sizes. Lime required for pH control ranged from 1.6 to 3.3 kg/t.

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Gravity Concentration Tests

A scoping gravity concentration test was conducted on a 12 kg (24.5 pound) split from each of the seven composites, at 80% -212 microns (65-mesh), to evaluate response to gravity concentration and to generate gravity tailing for flotation testing.

The composites responded moderately well to gravity concentration. Gold values reporting to a gravity cleaner concentrate with an average mass of 0.09% of the feed and an average grade of 146 g Au/t, represented gold recoveries of between 19.4% and 65.3% (38.0% average). Gravity gold recovery was not correlated to feed grade. These results indicate moderate potential for gold recovery by gravity concentration. Silver recoveries were much lower and did not exceed 11% to the cleaner concentrate.

Flotation Concentration Tests

Bulk sulfide flotation tests were conducted on each of the seven non-oxide composites, at an 80% -212 microns (65 mesh) feeds size, both with and without head-end gravity concentration treatment. Based on results from those tests, two master composites (4471-067 and 068) were prepared and used along with one of the original composites (4471-061) for more detailed testing. That testing included a series of grind optimization flotation tests on gravity tailing and a comparative whole ore flotation test at the indicated optimum grind size (80% -150 microns) (100 mesh). Based on results from these tests an overall master composite (4471-069) was prepared for bulk flotation, to generate concentrate for cyanidation testing.

A grind optimization test series (gravity/flotation) was conducted on two master composites (4471-067 and 4471-068) and on the original composite that gave below trend gold recovery at -212 microns (4471-061). Feed sizes evaluated were 80% -212 microns, -150 microns, -106 microns (150 mesh) and -75 microns (200 mesh). The -212 microns data was calculated for composites 4471-067 and 068, from testing conducted on the constituent composites comprising the two master composites. Master composite flotation feeds for the finer grind sizes were prepared by compositing gravity tailing generated from the constituent composites (80% -212 microns gravity feed size) before regrinding to the required sizes. Comparative whole ore (without gravity concentration) flotation tests were conducted on the three composites at the -212 microns and -106 microns feed sizes. Master composite -212 microns feed size data were again calculated from testing on the constituent composites.

Summary results from the flotation variability, grind optimization and bulk flotation tests, with and without gravity concentration, are presented in Appendix B Table 11 and Appendix B Table 12.

The flotation variability testing showed that the non-oxide composites generally responded well to direct flotation treatment, at an 80%-212 microns feed size. Gold values reporting to the flotation rougher concentrates (5.3% mass pull average) generally ranged from 72% to 95% and averaged 78.0% of the total gold contained in the feed.

Composite 4471-061 gave a significantly below trend gold recovery (51.9%) at the 212 microns feed size. Omitting that composite improved the average gold recovery to 82.3%. This composite was comprised of relatively deep (112 – 158 meters) (367 – 517 feet) generally tuff lithology material from hole IFM19_064 drilled from the north-central zone of the Main area. This composite appeared to be more sensitive to grind size than the others and gave higher recoveries at finer grind sizes.

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Direct flotation silver recoveries obtained at the 212 microns feed size were significantly lower and ranged from 23.7% to 66.4%. Sulfide sulfur recoveries to the rougher concentrate (measured only for the tests without gravity pre-concentration) were consistently high and averaged 90%.

Removing a gravity cleaner concentrate before flotation improved average gold and silver recoveries obtained at the 212 microns feed size to 84.9% and 69.1%, respectively. The feeds for these tests were the recombined gravity (cleaner plus rougher) tailing generated during the gravity testing described in the preceding section. As shown in Figures FMM-5 and 6, flotation rougher tail grades generally were incrementally lower when the gravity concentrate was removed from the flotation feed, but in most cases the difference fell within experimental and assay precision limits.

Flotation grind optimization test results showed that grinding the gravity tailing from 80%-212 microns to 80% -150 microns (100 mesh) in size before flotation significantly improved gold recovery from composite 4471-061 and significantly improved flotation silver recovery from the two other composites. Grinding finer than 80%-150 microns in size was not effective in significantly improving recoveries. Gold recoveries to the combined (gravity and flotation) concentrates generated at the 150 microns feed sizes were 83.9% (Comp. 4471-061), 90.7% (4471-067) and 88.7% (4471-068). Respective silver recoveries were 82.7%, 60.9% and 88.3%. As shown in Figures FMM-6 and 7, flotation tail grades in some cases were decreased significantly by grinding to 150 microns in size but were very similar at 150 microns and finer grind sizes.

Comparative flotation tests conducted without gravity concentration, at the 212 microns (65 mesh) and 150 microns feed sizes indicated no significant benefit to gravity pre-concentration at the 150 microns feed size. Some variability in head grade, which caused some variability in recoveries, was noted for the whole ore flotation tests. Comparison of the flotation tail grades (with and without gravity pre-concentration) was believed to provide a more accurate assessment of grind sensitivity.

Five bulk (8 kg) direct (without gravity pre-concentration) flotation tests were conducted on an overall master composite (4771-069), at an 80% -150 microns feed size, to generate concentrate for cyanidation testing. Results from those flotation tests are also presented in Appendix B Table 11 and Appendix B Table 12.

The flotation rougher concentrates produced (3.8% mass pull average) contained 81.2% to 89.7% (86.5% avg.) of the total gold and 69.3% to 74.9% (72.4% avg) of the total silver. Average rougher tail grade from these tests was 0.03 g Au/t and 2.7 g Ag/t. Concentrates from these flotation tests were used as feed for the concentrate regrind/cyanidation tests. Calculated head grades from those cyanidation tests were used as the concentrate grade for these tests. Sulfide sulfur recoveries were not tracked for these flotation tests.

In summary, flotation test results indicated that a feed size of 80%-150 microns was optimum for maximizing flotation gold and silver recovery from the Florida Mountain non-oxide mineralization. Gravity pre-concentration resulted in negligible improvement of gold and silver recoveries to the concentrate at this feed size.

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Flotation Product Cyanidation Tests

Agitated cyanidation tests were conducted on flotation rougher concentrate, generated from the bulk flotation tests described in the preceding section, to optimize regrind size for cyanide leaching of the concentrate. Regrind sizes ranging from 80% -39 microns to 80% -14 microns were tested. Cyanidation test conditions and results are shown in Table 13.9, Figure 13.7 and Figure 13.8.

Table 13.9 Florida Mountain Flotation Concentrate Regrind/Cyanidation

(Non-Oxide Master Composite 4471-069)

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Figure 13.7 Recovery vs. Regrind, Flotation Conc. Cyanidation, Florida Mountain

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Figure 13.8 Gold and Silver Recovery Rates, Flotation Ro. Concentrate, Florida Mountain

(Comp. 4471-069, 80%-20 microns Regrind)

Results showed that gold recovery increased incrementally with decreasing regrind size from 85.6% of gold contained in the concentrate (80% -39m) to 89.5% (80% -14 microns). Silver recovery was more sensitive to regrind size and increased from 75.0% (80% -30 microns) to 87.2% (80% -20 microns) of silver contained in the concentrate. Grinding from 20 microns to 14 microns did not further increase silver recovery.

Based on the average flotation recoveries (86.5% Au and 72.4% Ag) to the bulk rougher concentrates used for these cyanidation tests, the concentrate leach recoveries represented overall (flotation/ concentrate leach) recoveries between 74.0% and 77.7% gold and between 54.3% and 63.2% silver. Considering the flotation testing on the individual composites discussed in this section, higher flotation recoveries are expected from these materials in a continuous flotation circuit.

Reagent consumptions were relatively low and equivalent to 0.30 kg NaCN/t and 0.1 kg/t on a “whole ore” basis, at the 20 microns regrind size.

Select flotation rougher tails samples (Comps. 4471-061, 067 and 068), generated from gravity tailing at a 150 microns feed size, were used as feed for agitated cyanidation tests to evaluate the potential for recovery of gold and silver reporting to the flotation tailing by leaching.

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Cyanidation gold recoveries obtained from the flotation tailing ranged from 33.3% to 50.0%, which was equivalent to between 6% and 10% of gold contained in the whole ore (gravity feed). Silver recoveries obtained from the flotation tailing ranged from 53.8% to 70.0%, which was equivalent to between 8% and 22% of silver contained in the whole ore. Flotation tailing leach cyanide consumption was equivalent to between 0.19 and 0.28 kg NaCN/t whole ore and lime requirement was equivalent to between 1.5 and 1.9 kg CaO/t whole ore. Given the low-grade nature of the flotation tailing, it is unlikely that leaching the flotation tailing would be economical.

Solid Liquid Separation Testing – Florida Mountain Flotation Tailing

Flotation tailing from two of the bulk (8 kg) flotation tests used to generate concentrate for cyanide leach tests were shipped from McClelland to Pocock Industrial, in Salt Lake City, Utah for solid liquid separation (“SLS”) testing. The results were summarized by Pocock (2021). The overall objective for the testing was to develop a general set of data for design of thickening and filtration equipment intended to dewater the Florida Mountain flotation tailing for final disposal. Testing included flocculant screening tests, static and dynamic settling tests, slurry rheology tests, vacuum filtration tests and pressure filtration tests.

Sample characterization showed that the average measured P80 of the tailing was 129 microns. This was noted to be somewhat finer than measured by McClelland for the flotation test feeds (150 microns) (100 mesh). The flotation tailing was also noted by Pocock to have a relatively high fines content (47% -25 microns).

Flocculant screening testing showed that most anionic flocculants provided some degree of response for the samples, but that poor liquor clarity was almost universally a problem. Addition of lime and use of cationic flocculant did not significantly improve clarity. Only SNF AN970SH (a medium to high molecular weight, 70% charge density, anionic polyacrylamide) flocculant was shown to produce reasonable clarity while maintaining robust flocculant structure and rapid settling rates. This flocculant was chosen for further test work. Static thickening tests were conducted, and recommended design parameters based on those results were reported.

It was noted that the samples flocculated reasonably well and exhibited settling performance at feed solids concentrations above 15% to 20%. Required SNF AN970SH flocculant dosage was reported to be 65 g/t. Conventional thickener unit areas ranged from 0.159 to 0.202 m2/tpd. These unit areas were based on an underflow density of 48 percent. Although not achieved in static testing, it was noted that underflow densities could exceed 51 percent based on rheology.

Dynamic thickening tests were conducted to determine recommended maximum hydraulic design basis for high-rate thickener design. Resulting recommended design criteria and operating parameters were reported (Pocock, 2021). Suggested maximum design hydraulic loading rate was reported as 2.03 to 2.22 m3/m2hr at a maximum feed solids concentration of 15% to 17.5%. Predicted underflow density was 51% to 53.9%.

Vacuum filtration tests were conducted on the flotation tailing. It was reported that “Vacuum filtration was unable to produce stackable cakes for either material even with relatively long dry times.” (Pocock, 2021). The tailing was not a good candidate for vacuum filtration. It was noted that the production rates achieved during testing were high enough to suggest that pressure filtration would likely be a reasonable flowsheet alternative.

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13.3.2.6Florida Mountain Comminution Testing

A total of nine non-oxide composites, including two from the PEA testing program and seven from the PFS testing program, were used for Bond ball mill work index determinations. Ball mill index testing was conducted using standard Bond methods. Abrasion index tests were also conducted, using a Bond impellor-tumbler impact abrasion test method, on the same seven PFS composites. Summary comminution test results are shown in Appendix B Table 13.

Bond abrasion indices were variable and classified as lightly abrasive to very abrasive. Bond ball mill work indices were variable and classified as medium to hard. Ball mill work indices ranged from 12.97 to 20.12 kilowatt-hour per metric ton.

13.4Recovery Models and Reagents

13.4.1Oxide and Mixed Materials Heap-leach Recovery and Reagent Estimates

Heap-leach gold and silver recovery estimates were developed based primarily on column-leach test data and to a lesser degree on bottle-roll tests from the PEA and PFS testing programs. Parameters were assigned for each of the areas tested. Where column test data were available, gold recoveries were based on discounted column test recoveries (generally discounted by 3% recovery). Silver recoveries were based on extrapolated column test silver recoveries, to account for the generally slower recovery rates and the incomplete leaching generally achieved during column testing. Where column test data were not available, heap-leach recoveries were estimated using available bottle-roll test data and the correlations established between column and bottle-roll test recovery results. For areas where neither column, nor bottle-roll test data were available (DeLamar/DeLamar North area mixed material, and Ohio area mixed material), average recovery data from select other areas were used.

Cyanide consumptions were based on discounted cyanide consumptions from 60-day column-leach cycles (3:1 factor was applied). Lime consumption for the oxide material was based on column test lime additions. Lime consumption for the mixed materials was based on column test lime additions increased by a factor of 25%, to allow for mitigation of potential long-term acid generation from partial oxidation/weathering of the sulfide minerals contained in those materials.

Based on load/permeability test data and the heap stacking plan, it was estimated that 45% of the DeLamar oxide and mixed material will require agglomeration pretreatment with cement. For the purposes of reagent consumption estimates, that cement consumption was assigned on a weighted basis to all the DeLamar oxide and mixed materials. Lime consumptions were discounted accordingly on a 1:1 basis with the cement added. The Florida Mountain oxide and mixed feeds do not require agglomeration pretreatment.

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Estimated recoveries and reagent consumptions for heap leaching of the DeLamar and Florida Mountain oxide and mixed material types are shown in Table 13.10.

Table 13.10 PFS Heap-Leach Recovery and Reagent Consumption Estimates

13.4.2DeLamar Non-Oxide Mill Recovery and Reagent Estimates

DeLamar non-oxide feed will be sourced primarily from the Sullivan Gulch area, with lesser amounts from the other DeLamar areas. Cyanide leach amenability of the Sullivan Gulch non-oxide material is highly variable. A review by Integra of the CN/FA gold data from Integra 2018 drilling was conducted to evaluate the distribution of cyanide leach amenability throughout the Sullivan Gulch area. This included data from all Integra drill holes in DeLamar non-oxide material with both cyanide soluble gold and fire assay data available. This data set considered all non-oxide material with a grade above 0.3 g Au/t equivalent and came from 36 drill holes totaling 2,604 linear meters (8,542 feet) of drill samples. These samples had an average interval length of 1.58 meters (5.18 feet). The distribution of gold CN/FA ratio from the drill data is presented graphically in Appendix B Figure 5.

Extensive agitated cyanidation testing was conducted on both “whole ore” and on reground flotation concentrate, generated from Sullivan Gulch non-oxide metallurgical composites, at a VFG size (nominally 20 microns). To consider both data sets together, gold recoveries from the “whole ore” leach tests were discounted to account for estimated flotation gold recovery to calculate estimated flotation concentrate regrind/cyanidation gold recoveries. The two data sets (VFG-leach and flotation concentrate cyanidation gold recoveries) were plotted versus CN/FA ratio for the corresponding metallurgical composites, as shown in Appendix B Figure 6.

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The correlation between estimated flotation – concentrate cyanidation gold recovery (with a maximum recovery set at 60%) and CN/FA ratio was applied to the Sullivan Gulch non-oxide drill interval CN/FA data set (Appendix B Figure 5) on a drill interval basis, to estimate mill (flotation – concentrate regrind/cyanidation) gold recovery from the Sullivan Gulch non-oxide material.

Sullivan Gulch silver recoveries by cyanidation after very fine grinding (either “whole ore” or flotation concentrate) were much less variable. Results from the concentrate cyanidation and “whole ore” cyanidation tests (factored for estimated flotation recovery) were considered for estimating mill silver recovery for the Sullivan Gulch non-oxide materials.

In the case of Glen Silver and the other DeLamar areas, a smaller number of samples were tested. Results from the available flotation with flotation concentrate cyanidation tests were used, along with results from the whole ore VFG/cyanidation tests conducted on composites from those areas to estimate the respective mill gold and silver recoveries.

Estimated mill recoveries and reagent consumptions are presented in Table 13.11 and Table 13.12.

Table 13.11 DeLamar Non-Oxide Mill Recoveries

DLM is DeLamar

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Table 13.12 DeLamar Non-Oxide Reagent Estimates

13.4.3Florida Mountain Non-Oxide Mill Recovery and Reagent Estimates

Grade versus recovery relationships were established for recoveries of gold and silver to the flotation concentrate produced from the Florida Mountain non-oxide material, at an 80% -150 microns (100 mesh) grind size. These relationships were developed based on flotation testing conducted during the PEA and PFS metallurgical testing programs. The modeled flotation grade-recovery relationships for gold and silver are shown, along with results from representative flotation tests, in Appendix B Figure 7 and Appendix B Figure 8, respectively. Maximum flotation gold and silver recoveries were set at 96% and 89%, respectively, based on the available flotation test results.

Flotation concentrate regrind/cyanidation gold and silver recoveries were estimated at 91% and 88%, respectively of gold and silver contained in the flotation concentrate.

Based on the flotation and concentrate regrind/cyanidation gold and silver recoveries discussed in the preceding paragraphs, estimated overall recoveries and reagent consumptions are shown in Table 13.13 and Table 13.14, respectively. Maximum overall (flotation/concentrate regrind-cyanidation) gold recovery is estimated at 87%. Maximum overall silver recovery is estimated at 77%.

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Table 13.13 Florida Mountain Non-Oxide Overall Mill Recoveries

Table 13.14 Florida Mountain Non-Oxide Overall Reagent Estimates

13.5Summary Statement

Mr. McPartland has reviewed the historical metallurgical studies and the metallurgical studies conducted during 2018 through 2021 and concludes that the samples used during the 2018 through 2021 metallurgical studies are reasonably representative considering both the stage of the project development and the magnitude of the testing completed as of the effective date of this report. However, further testwork of samples collected from portions of the deposit, particularly those displaying high degrees of variability in metallurgical response, will be needed as the project advances. Other than as discussed herein, the author is not aware of any processing factors or deleterious elements that could have a significant effect on the potential economic extraction.

The oxide and mixed materials from both the DeLamar and Florida Mountain deposits can be processed by heap-leach cyanidation. Oxide materials from both deposits are expected to give relatively high (75% – 89%) heap-leach gold recoveries, which are not particularly sensitive to feed size. Mixed materials from both deposits give lower and more variable gold recoveries (45% – 77% at a 12.7mm crush size) (0.5-inch) which are more sensitive to feed size. Heap-leach silver recoveries expected from the oxide and mixed materials are significantly lower, more variable (11% – 74% at a 12.7mm crush size), and more sensitive to feed size. A relatively fine (80% -12.7mm) (-0.5-inch) feed size was selected for heap leaching the DeLamar and Florida Mountain oxide and mixed materials to maximize silver recovery, and in the case of the mixed materials, to maximize gold recovery. Some of the DeLamar oxide and mixed materials contain elevated levels of clay and will require agglomeration pretreatment before heap leaching. None of the Florida Mountain materials require agglomeration. Non-oxide materials from both deposits generally are not expected to be amenable to heap-leach cyanidation.

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Non-oxide materials from the DeLamar deposit give generally low but highly variable cyanidation gold recoveries, even after very fine grinding. These materials contain significant amounts of gold that are locked in sulfide mineral particles. For these materials, oxidative pretreatment (such as the Albion process) of sulfide minerals is required for liberation of gold before high cyanidation recoveries can be obtained. In contrast, silver contained in the DeLamar non-oxide materials generally can be liberated to a large degree by very fine grinding (-20 microns). Cyanidation silver recoveries after very fine grinding generally are relatively high (approximately 80%). The DeLamar non-oxide materials respond well to flotation at a moderate grind size (150 microns) (100 mesh). The resulting flotation concentrate responds well to cyanide leaching after very fine regrinding for recovery of contained silver. Some gold is also recovered by cyanide leaching of the reground flotation concentrate, but those recoveries generally are low.

Non-oxide materials from the Florida Mountain deposit are amenable to cyanidation treatment at moderately fine grind sizes (75 microns and finer) (200 mesh and finer) and give relatively high gold recoveries (generally >80%) and moderate silver recoveries (generally >50%). These recoveries can be improved by finer grinding. These materials respond well to upgrading by flotation at a moderate grind size (150 microns) and cyanidation gold and silver recoveries from the resulting concentrates can be maximized by very fine regrinding (20 microns). In contrast to the DeLamar non-oxide materials, oxidative pretreatment of contained sulfide minerals is not required to achieve high cyanidation gold recoveries from the Florida Mountain non-oxide feeds.

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14.0MINERAL RESOURCE ESTIMATES

14.1Introduction

The updated mineral resource estimations for the DeLamar project, which include resources of the DeLamar and Florida Mountain areas, were completed for public disclosure in accordance with the guidelines of NI 43-101. The mineral resources were estimated under the supervision of Mr. Gustin, a qualified person with respect to mineral resource estimations under NI 43-101. Mr. Gustin is independent of Integra by the definitions and criteria set forth in NI 43-101; there is no affiliation between Mr. Gustin and Integra except that of independent consultant/client relationships.

This report presents updated gold and silver resources for the DeLamar and Florida Mountain deposits that have an effective date of March 1, 2021, the date the latest assays were received by MDA from Integra; drilling has continued almost unabated at the project since that time.

The DeLamar project resources are classified in order of increasing geological and quantitative confidence into Inferred, Indicated, and Measured categories in accordance with the “CIM Definition Standards – For Mineral Resources and Mineral Reserves” (2014) and therefore NI 43-101. CIM mineral resource definitions are given below, with CIM’s explanatory text shown in italics:

Mineral Resource

Mineral Resources are sub-divided, in order of increasing geological confidence, into Inferred, Indicated and Measured categories. An Inferred Mineral Resource has a lower level of confidence than that applied to an Indicated Mineral Resource. An Indicated Mineral Resource has a higher level of confidence than an Inferred Mineral Resource but has a lower level of confidence than a Measured Mineral Resource.

A Mineral Resource is a concentration or occurrence of solid material of economic interest in or on the Earth’s crust in such form, grade or quality and quantity that there are reasonable prospects for eventual economic extraction. The location, quantity, grade or quality, continuity and other geological characteristics of a Mineral Resource are known, estimated or interpreted from specific geological evidence and knowledge, including sampling.

Material of economic interest refers to diamonds, natural solid inorganic material, or natural solid fossilized organic material including base and precious metals, coal, and industrial minerals.

The term Mineral Resource covers mineralization and natural material of intrinsic economic interest which has been identified and estimated through exploration and sampling and within which Mineral Reserves may subsequently be defined by the consideration and application of Modifying Factors. The phrase ‘reasonable prospects for eventual economic extraction’ implies a judgment by the Qualified Person in respect of the technical and economic factors likely to influence the prospect of economic extraction. The Qualified Person should consider and clearly state the basis for determining that the material has reasonable prospects for eventual economic extraction. Assumptions should include estimates of cutoff grade and geological continuity at the selected cut-off, metallurgical recovery, smelter payments, commodity price or product value, mining and processing method and mining, processing and general and administrative costs. The Qualified Person should state if the assessment is based on any direct evidence and testing.

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Interpretation of the word ‘eventual’ in this context may vary depending on the commodity or mineral involved. For example, for some coal, iron, potash deposits and other bulk minerals or commodities, it may be reasonable to envisage ‘eventual economic extraction’ as covering time periods in excess of 50 years. However, for many gold deposits, application of the concept would normally be restricted to perhaps 10 to 15 years, and frequently to much shorter periods of time.

Inferred Mineral Resource

An Inferred Mineral Resource is that part of a Mineral Resource for which quantity and grade or quality are estimated on the basis of limited geological evidence and sampling. Geological evidence is sufficient to imply but not verify geological and grade or quality continuity. 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.

An Inferred Mineral Resource is based on limited information and sampling gathered through appropriate sampling techniques from locations such as outcrops, trenches, pits, workings and drill holes. Inferred Mineral Resources must not be included in the economic analysis, production schedules, or estimated mine life in publicly disclosed Pre-Feasibility or Feasibility Studies, or in the Life of Mine plans and cash flow models of developed mines. Inferred Mineral Resources can only be used in economic studies as provided under NI 43-101.

There may be circumstances, where appropriate sampling, testing, and other measurements are sufficient to demonstrate data integrity, geological and grade/quality continuity of a Measured or Indicated Mineral Resource, however, quality assurance and quality control, or other information may not meet all industry norms for the disclosure of an Indicated or Measured Mineral Resource. Under these circumstances, it may be reasonable for the Qualified Person to report an Inferred Mineral Resource if the Qualified Person has taken steps to verify the information meets the requirements of an Inferred Mineral Resource.

Indicated Mineral Resource

An Indicated Mineral Resource is that part of a Mineral Resource for which quantity, grade or quality, densities, shape and physical characteristics are estimated with sufficient confidence to allow the application of Modifying Factors in sufficient detail to support mine planning and evaluation of the economic viability of the deposit. Geological evidence is derived from adequately detailed and reliable exploration, sampling and testing and is sufficient to assume geological and grade or quality continuity between points of observation. An Indicated Mineral Resource has a lower level of confidence than that applying to a Measured Mineral Resource and may only be converted to a Probable Mineral Reserve.

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Mineralization may be classified as an Indicated Mineral Resource by the Qualified Person when the nature, quality, quantity and distribution of data are such as to allow confident interpretation of the geological framework and to reasonably assume the continuity of mineralization. The Qualified Person must recognize the importance of the Indicated Mineral Resource category to the advancement of the feasibility of the project. An Indicated Mineral Resource estimate is of sufficient quality to support a Pre-Feasibility Study which can serve as the basis for major development decisions.

Measured Mineral Resource

A Measured Mineral Resource is that part of a Mineral Resource for which quantity, grade or quality, densities, shape, and physical characteristics are estimated with confidence sufficient to allow the application of Modifying Factors to support detailed mine planning and final evaluation of the economic viability of the deposit. A Measured Mineral Resource has a higher level of confidence than that applying to either an Indicated Mineral Resource or an Inferred Mineral Resource. It may be converted to a Proven Mineral Reserve or to a Probable Mineral Reserve.

Mineralization or other natural material of economic interest may be classified as a Measured Mineral Resource by the Qualified Person when the nature, quality, quantity and distribution of data are such that the tonnage and grade or quality of the mineralization can be estimated to within close limits and that variation from the estimate would not significantly affect potential economic viability of the deposit. This category requires a high level of confidence in, and understanding of, the geology and controls of the mineral deposit.

Modifying Factors

Modifying Factors are considerations used to convert Mineral Resources to Mineral Reserves. These include, but are not restricted to, mining, processing, metallurgical, infrastructure, economic, marketing, legal, environmental, social and governmental factors.

14.2DeLamar Project Data

The DeLamar project gold and silver resources were estimated using drill data generated by Integra as well as the data derived from the exploration programs of the various historical operators discussed in Section 10.0. This information, most importantly including the data derived from RC, conventional rotary, and diamond-core drill holes, current topography, historical documentation of the as-mined open-pit topographies, cross-sectional lithological and structural interpretations, and documentation of historical underground workings, were provided to MDA by Integra.

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14.2.1Drill-Hole Data

The historical project data utilized mine-grid coordinates, a local grid system in Imperial units developed in the early 1970s and used throughout the life of the DeLamar project open-pit mining operations. The original down-hole drill intervals were in feet, and the gold and silver analyses were primarily reported in ounces per ton. In 2018, Integra completed a LiDAR aerial survey of the entire DeLamar project area, obtained historical survey data in both mine-grid and real-world coordinates, and transformed the drill-hole locations into UTM Zone 11 NAD 83 coordinates with the assistance of MDA. All project down-hole drill depths, assays, and geologic logging intervals were then converted into meters and grams-per-tonne.

As discussed in Section 10.7, drill intervals identified as having significant sample quality issues, including poor sample recoveries and down-hole contamination, were excluded from use in the resource estimation. In addition, sample intervals of colluvial materials were either explicitly excluded from the gold-and silver-domain modeling described below, as was commonly the case for the DeLamar deposit, or tagged for exclusion directly in the project databases, as was the case for both the DeLamar and Florida Mountain database. The excluded colluvial sample intervals are commonly mineralized, especially at Florida Mountain where significantly mineralized colluvium was frequently intersected in the top few meters of drill holes.

After verifying the historical data, MDA constructed resource databases for the DeLamar and Florida Mountain areas. Gold and silver values used in the current resource estimations were prioritized as follows: fire assays by outside labs were given top priority, followed by fire assays by the on-site mine lab, with mine-lab AA gold analyses selected only in cases where no other data were available. Certain low-precision gold analyses and all mine-lab AA silver assays were identified, flagged in the resource databases, and not used in the estimation of the project resources (see Section 12.0).

14.2.2Topography

Integra provided MDA with project-wide elevation data from their LiDAR survey, which was used to create digital topographic surfaces for both the DeLamar and Florida Mountain deposit areas. These current topographic surfaces reflect post-mining reclamation, including re-contouring of waste dumps and the partial backfilling of many of the open pits.

Integra also provided MDA with original historical paper plots of final post-mining topographies of the historical open pits at both the Florida Mountain and DeLamar areas. MDA used these paper plan maps to create digital ‘as-mined’ topographic surfaces that encompasses the areas of historical open-pit mining. Based on other historical data, including blast-hole information, as well as the current topography derived from the LiDAR survey and Integra drilling through backfilled areas, Mr. Gustin believes the modeled as-mined surfaces reasonably represent the volumes mined during the historical open-pit operations.

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14.2.3Modeling of Historical Underground Workings

Integra provided MDA with three-dimensional digital linework created by Kinross that represents historical drifts, crosscuts, and developmental workings in the DeLamar area. This modeling by Kinross, which was based on historical records reviewed by the authors, indicates that the historical underground workings in the DeLamar area lie almost entirely inside of the historical North DeLamar and Sommercamp open pits. However, the drifts along the mined vein structures and related developmental winzes were useful in the modeling of the unmined gold and silver resources lying below and adjacent to the pits, as they provided evidence of the strikes and dips of the mined mineralized structures.

Underground workings at Florida Mountain, including drifts, cross cuts, winzes, shafts, and stopes, are documented by a series of original hand-drafted level plans, long sections, and cross sections in the possession of Integra that date from the late 1800s to the early 1900s. MDA used these drawings to create three-dimensional digital models of the underground workings and stopes, although there is little information as to the widths of the stopes. While these drawings are unlikely to include all historical underground mining that took place at Florida Mountain, there is good evidence that a high percentage of the stopes from the Black Jack – Trade Dollar workings are represented.

14.3Geological Modeling

Integra completed digital lithological and structural interpretations on sets of cross sections that span the extents of the Florida Mountain and DeLamar resource areas. These cross sections were used as the base for modeling the gold and silver mineral domains discussed in Section 14.8.1. Lithological contacts that influenced the distributions of the gold and silver mineralization, as well as faults modeled on sections by Integra and high-angle mineralized zones modeled by MDA, were represented as three-dimensional wireframe surfaces that served as guides for the detailed modeling of the gold and silver mineralization as part of the estimation of the project mineral resources.

14.4Deposit Geology Pertinent to Resource Modeling

The DeLamar area mineralization is predominantly influenced by moderate- to high-angle zones of higher-grade mineralization and associated much larger bodies that halo the higher-grades. In some areas, moderately dipping mineralization flattens upwards to the northeast, although significant portions of these more shallowly dipping zones of mineralization were mined in the historical operations. The mineralization is overwhelmingly hosted in the felsic volcanic units that lie above the lower basalt and below the banded rhyolite. While only minor mineralization has been drilled within the lower basalt, low-grade mineralization does occur locally within banded rhyolite that lies in the uppermost portions of the felsic package.

At Florida Mountain, the gold and silver mineralization drilled to date also occurs primarily within felsic volcanic units, which in this area overlie Cretaceous granodiorite. The granodiorite hosts most of the high-grade veins that were the focus of the historical underground mining at Florida Mountain, although the Trade Dollar vein in particular was mined into the felsic package, locally through to the present-day surface. The Florida Mountain mineralization that comprises the current resources occurs along multiple, broad, north- to north-northwest-striking zones straddling the contacts of rhyolitic intrusive necks that intrude the felsic volcanic units. From west to east, these zones are centered on the historical Ontario, Tip Top, Arcuate, Alpine, Stone Cabin, and Trade Dollar-Black Jack mining and exploration areas. In detail, each of these mineralized zones are comprised of complex networks of thin, interweaving mineralization that forms what can be considered large-scale stockwork zones, and taken as a whole, these zones formed bulk-mineable bodies. These zones blossom outwards from the intrusive rhyolite necks upwards towards the surface and collapse inward towards the rhyolite necks at deeper levels in the felsic volcanic package. The continuations of the mineralization into higher-grade, more discrete zones at depth, especially into the granodiorite, are presently being explored as potential underground mineable targets.

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Mr. Gustin reviewed the distribution of the silver mineralization intersected in drilling carefully, especially in the silver-rich DeLamar area, to discern the presence (or absence) of potential supergene-enriched zones that would be relevant to the resource modeling. Only a few limited areas were found that are suggestive of possible supergene enrichment, but the evidence is not conclusive. Several historical reports state that secondary enrichment of silver probably occurred on a limited scale, although the evidence cited is restricted to the presence of cerargyrite.

14.5Water Table

The 1974 historical feasibility study, which focused on the Sommercamp and North DeLamar areas, stated that surface oxidation generally does not extend deeper than about 55 meters (180 feet) from the surface, except along fault zones (Earth Resources Company, 1974). The water table was stated in the 1974 study to lie at an elevation of approximately 1,845 meters (~6,053 feet), considerably deeper than the level of oxidation. These statements presumably applied only to the two deposit areas that were the subject of the historical feasibility study. A later mine document reported a water table depth of 1,810 meters (~5,938 feet) at the north end of the Sommercamp – Regan zone, which at the time included what was referred to as South Wahl (Pancoast, 1990). Ms. Richardson indicated to MDA that the water table lies near the bottoms of the North DeLamar and South Wahl pits, at elevations of about 1,820 and 1905 meters (5,971 and 6,250 feet), respectively.

14.6Oxidation Modeling

Integra completed comprehensive logging of oxidation of the historical RC and rotary holes using the chipboards present at the project site. This information was then combined with oxidation logging of Integra RC chips and drill core and added to the project databases. While earlier resource modeling relied almost exclusively on these visual logging codes, there is now sufficient Integra drilling to incorporate chemical analyses pertinent to oxidation state into the modeling. The chemical data include cyanide-leach analyses of drill-sample pulps, ICP sulfur data, and limited LECO sulfur speciation data. The cyanide-leach analyses were used to calculate cyanide-leach-gold-to-fire-assay-gold (“CN/FA”) ratios.

MDA created wireframe solids of oxide and non-oxide zones based on the visual logging of oxidation state and the chemical data summarized above, and the resultant solids were used to code the DeLamar and Florida Mountain block models. Model blocks lying between those coded as oxide or non-oxide, which are comprised of oxide, partially oxidized, and non-oxide materials that lack continuity to be modeled separately, were assigned a “mixed” code.

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Despite the addition of the chemical data with every new hole drilled by Integra, the visual logging codes are still the dominant input to the modeling of oxidation state. In order to evaluate the accuracy of the logging, MDA requested Integra to send a large batch of Integra drill-sample pulps for cyanide-leach assay; these sample pulps were from holes drilled prior to the now routine cyanide-leach analyses of all samples within mineralized zones, and they therefore lacked cyanide-leach assays. The incorporation of this new cyanide-leach data led, in general, to small expansions of the previously modeled mixed zones at the expense of previously modeled non-oxide zones. While the impacts were small, they indicate that the modeling of non-oxide materials at the mixed / non-oxide boundary will tend to slightly overstate the non-oxide materials in areas lacking chemical data pertinent to the assignment of oxidation state.

The coding of oxidation state in the DeLamar and Florida Mountain models determines the potential processing option for each block in each model, and therefore the resource cutoff grade that applies to each block (discussed further below).

The CN/FA ratios indicate that there are areas of mixed materials within the modeled non-oxide zone at the Sullivan Gulch portion of the DeLamar area, but there are insufficient CN/FA ratio assays to model these zones confidently. These unmodeled mixed materials likely are related to fault zones through which oxygenated meteoric waters percolated and partially oxidized otherwise non-oxide zones.

14.7Density Modeling

A number of references to density values were reviewed in the available historical records, including some density studies with limited numbers of actual density determinations listed. These datasets are generally only partially documented, with many lacking a description of the determination methods. While the methodologies used for the density determinations are often unclear, the records indicate determinations were done by a variety of methods, including water displacement, water immersion, volume/weight, and nuclear methods.

MDA compiled the data from two of the more completely documented historical specific gravity (“SG”) studies of samples from the DeLamar area. The 13 measurements yield an average SG of 2.31. A total of 12 historical SG determinations from Florida Mountain drill core compiled by MDA average 2.41. Historical DeLamar and Florida Mountain resource and reserve estimations undertaken during the open-pit operations most commonly used a global tonnage factor (mineralized and unmineralized rock) of 13.5 ft3/ton, which equates to an SG of 2.37. The historical open-pit operation used a wet density of 13.5 ft3/ton throughout the life of the mine to determine mill-feed tonnages and waste. Based on various measurements, the mine assumed 7.5% moisture in the mined materials at DeLamar and 6% at Florida Mountain. These values equate to global (dry) SGs of 2.21 for DeLamar and 2.24 for Florida Mountain.

Integra routinely measured the SG of selected samples of its drill core using the water immersion method. Table 14.1 and Table 14.2 summarize Integra’s SG results for samples of core from holes drilled at the DeLamar and Florida Mountain deposit areas, respectively, as well as the SGs assigned to the model blocks. The results are compiled by oxidation state and whether the samples are within the gold and/or silver mineral domains that constrain the resource estimations or lie outside of the domains (“Mineralized” or “Unmineralized”, respectively in Table 14.1 and Table 14.2).

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Table 14.1 Integra Specific Gravity Determinations from DeLamar Deposit Drill Core

Table 14.2 Integra Specific Gravity Determinations from Florida Mountain Deposit Drill Core

The data for both resource areas show increasing SG values from oxide to mixed to non-oxide zones, which is at least in part due to decreasing effects of weathering (oxidation). An examination of the raw data also indicates a suggestion that SG may increase slightly as metal grades increase, but the data quantity is insufficient to be confident that this is the case.

Specific gravity values of 1.82 and 1.88 were assigned to backfill and dump materials at the DeLamar and Florida Mountain deposits, respectively.

14.8DeLamar Area Gold and Silver Modeling

14.8.1Mineral Domains

A mineral domain encompasses a volume of rock that ideally is characterized by a single, natural grade population of a metal that occurs within a specific geologic environment. In order to define the mineral domains at the DeLamar project, the natural gold and silver populations were first identified on population-distribution graphs that plot the gold-grade and silver-grade distributions of all drill-hole assays at, in this case, the DeLamar area. This analysis led to the identification of low-, medium-, and high-grade populations for both gold and silver. Ideally, each of these populations can then be correlated with specific geologic characteristics that are captured in the project database, which can be used in conjunction with the grade populations to interpret the bounds of each of the gold and silver mineral domains. The approximate grade ranges of the low-grade (domain 100), medium-grade (domain 200), and higher-grade (domain 300) domains are listed in Table 14.3.

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Table 14.3 Approximate Grade Ranges of DeLamar Area Gold and Silver Domains

Domain g Au/t g Ag/t
100 ~0.15 to ~1 ~5 to ~30
200 ~1 to ~6 ~30 to ~200
300 > ~6 > ~200

The DeLamar gold and silver mineral domains were modeled by interpreting silver, followed by gold, polygons on a set of vertical, 30-meter (98.4-foot) spaced, northwest-looking (Az. 320°) cross sections that span the presently drilled extents of the deposit. The mineral domains were interpreted using the gold and silver drill-hole assay data, associated drill-hole lithologic codes, documented descriptions of the mineralization, the historical underground workings, and Integra’s geological cross sections.

The mineral-domain modeling for the current resource estimation was aided to a significant extent by the lithologic, structural, and mineralization cross sections completed by Integra. The Integra cross sections, coupled with the closely spaced drilling throughout the resource area, were critical to the high-confidence modeling of the mineral domains.

The low-grade gold and silver domains (“100” domains) generally encompass relatively extensive southwest-dipping bodies lying within the various felsic volcanic units that lie between the banded rhyolite and the lower basalt. Relatively restricted, moderately to steeply dipping zones of mineralization in the mid-grade and higher-grade domains (domains “200” and “300”, respectively) occur within the broad extents of the lower-grade mineralization.

In areas where all volcanic units from the lower basalt to the flow-banded rhyolite were preserved, the higher-angle mineralization often bends into approximate parallelism with the basal contact of the flow-banded rhyolite. The domain 200 and 300 mineralization then extends laterally in a northeasterly direction along or close to this contract. The portions of the DeLamar deposit in which the high-grade domain occurs, both in the high-angle zones and, most importantly, the lower-angle zones below the flow-banded rhyolite, were preferentially mined during the historical open-pit operations. Therefore, few of such shallow occurrences of the low-angle high-grade zones remain. The most important example of significant mining in areas where the contact zone had been eroded is at the Sommercamp pit, where all but a few erosional remnants of the lower-angle mineralization were present prior to mining. In this case, the higher grades and frequency of the high-angle mineralization were sufficient to warrant its extraction.

The lower contact of the banded rhyolite, as well as the faults that are evidenced by its displacements, were modeled by Integra. This contact and the faults were used extensively in the mineral-domain modeling. Steeply dipping high-grade zones not associated with faults recognized by Integra were typically modeled by MDA as having steep southwesterly dips, which is consistent with historical underground stopes in the Sommercamp and North DeLamar areas.

The main DeLamar area mineralization, which includes the entire area of historical mining, extends continuously over a northwest strike extent of about three kilometers, a maximum northeast-southwest width of 1.2 kilometers (0.75 miles), and an elevation range of 570 meters (1,870 feet). The Milestone portion of the DeLamar mineralization, which lies about three-quarters of a kilometer (0.47 miles) northwest of the northwesternmost extents of the main DeLamar area, adds an additional 640 meters (2,100 feet) of strike to the resource modeling.

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Cross sections showing examples of the modeled gold and silver mineral domains for the Sommercamp – North DeLamar – Regan, and Glen Silver areas of the resources are shown in Figure 14.1 through Figure 14.4.

The final cross-sectional gold and silver mineral-domain polygons were projected horizontally to the drill data within each sectional window, and these three-dimensional polygons were then sliced horizontally at six-meter (19.68-foot) elevation intervals that match the mid-block elevations of the resource block model. The slices were used to create a new set of mineral-domain polygons for both gold and silver on level plans at six-meter (19.68-foot) vertical spacings. Level plans were used due to the predominance of moderately to steeply dipping mineralization, especially in medium- and high-grade domains.

Wireframe surfaces of faults, high-angle mineralized structures, and important lithologic contacts that focus or terminate mineralization were used to assist in the rectification of the mineral domains on long sections and level plans. The completed level-plan mineral-domain polygons serve to rectify the gold and silver domains to the drill-hole data at the scale of the block model.

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Figure 14.1 Cross Section 2190 NW Showing Sommercamp and N. DeLamar Gold Domains

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Figure 14.2 Cross Section 2010 NW Showing Sommercamp and N. DeLamar Silver Domains

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Figure 14.3 Cross Section 2790 NW Showing Gold Domains at Glen Silver

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Figure 14.4 Cross Section 2790 NW Showing Silver Domains at Glen Silver

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14.8.2Assay Coding, Capping, and Compositing

Drill-hole gold and silver assays were coded to the gold and silver mineral domains, respectively, using their respective cross-sectional polygons. Assay caps were determined by the inspection of population distribution plots of the coded assays grouped by domain to identify high-grade outliers that might be appropriate for capping. The plots were also evaluated for the possible presence of multiple grade populations within any of the domains. Descriptive statistics of the coded assays by domain and visual reviews of the spatial relationships of the possible outliers, and their potential impacts during grade interpolation, were also considered in the definition of the assay caps (shown in Table 14.4).

Each model block was coded to the volume percentage of each of the three modeled domains for both gold and silver, as discussed below. Volumes of blocks that were not entirely coded to the low-, mid-, and higher-grade mineral domains for either or both metals were assigned to domain “0” and were estimated using assays lying outside of the modeled domains. Table 14.4 shows the gold and silver assay cap applied to each of the domains.

Table 14.4 DeLamar Area Gold and Silver Assay Caps by Domain

Domain g Au/t No. of Samples Capped

(% of samples)
g Ag/t No. of Samples Capped

(% of samples)
0 2 17 (<1%) 160 3 (<1%)
100 2 55 (<1%) 125 17 (<1%)
200 6 17 (<1%) 200 52 (<1%)
300 40 5 (2.0%) 1,325 24 (1.3%)

Descriptive statistics of the capped and uncapped coded assays are provided in Table 14.5 and Table 14.6 for gold and silver, respectively.

Table 14.5 Descriptive Statistics of DeLamar Area Coded Gold Assays

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Table 14.6 Descriptive Statistics of DeLamar Area Coded Silver Assays

In addition to the assay caps, restrictions to the search distances of higher-grade composites within some of the domains were applied during grade interpolations (discussed further below). Search restrictions can minimize the number of samples subjected to capping while properly respecting the highest-grade populations within each domain.

The capped assays were composited at 3.05 meter (10-foot) down-hole intervals respecting the mineral domains. Descriptive statistics of DeLamar composites are shown in Table 14.7 and Table 14.8 for gold and silver, respectively.

Table 14.7 Descriptive Statistics of DeLamar Area Gold Composites

Table 14.8 Descriptive Statistics of DeLamar Area Silver Composites

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14.8.3Block Model Coding

The six-meter-spaced level-plan and long-sectional mineral-domain polygons were used to code a model comprised of 6 x 6 x 6-meter (19.68 x 19.68 x 19.68-foot) blocks. The model is rotated to a bearing of 320°, which is consistent with orientation of the cross sections. The percentage volume of each mineral domain (the “partial percentages”) for both gold and silver, as coded directly by the level plans, is stored within each block, as is the volume percentage of the block that lies outside of the modeled domains for both gold and silver.

Two topographic surfaces were used to code the block model: the as-mined and present-day surfaces discussed in Section 14.2.2. These digital topographic surfaces were used to define: (1) the percentage of each block that lies within bedrock; and (2) the percentage of each block that is comprised of backfill/dump material, which lies above the as-mined surface and below the present-day surface.

The modeled mineralization has a variety of orientations, which led to the construction of wireframe solids to encompass model areas with unique orientations of the mineralization. These solids were then used to code the model blocks to these specific areas.

The oxidation wire-frame solids described in Section 14.6 were used to code model blocks as oxide, mixed, or non-oxide.

Finally, the specific-gravity values discussed in Section 14.7 were assigned to the partial percentage of bedrock coded to each model block according to the coded oxidation state of the block; the partial percentage of each block coded to backfill/dump was assigned a density value of 1.82. These coded density values were then used in combination with the stored percentages of rock and fill to determine the tonnage of each model block.

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14.8.4Grade Interpolation

Parameters used in the estimation of gold and silver grades are summarized in Table 14.9.

Table 14.9 Summary of DeLamar Area Grade Estimation Parameters

Estimation Pass – Au + Ag Domain Search Ranges (meters) Composite Constraints
Major Semi-Major Minor Min Max Max/Hole
Pass 1 + 2 – Doman 100 60 60 20 2 12 4
Pass 1 + 2 – Doman 200 + 300 + 0 60 60 20 2 20 4
Pass 3 – Doman 100 60 60 60 1 12 4
Pass 3 – Doman 0 + 200 +300 170 170 170 1 20 4


Restrictions on Search Ranges
Domain Search Restriction Threshold Search Restriction Distance Estimation Pass
Au 100 >0.7 g Au/t 40 meters 1, 2
Au 300 >20 g Au/t 35 meters 1, 2, 3
Au 0 >0.5 g Au/t 6 meters 1, 2, 3
Ag 300 >400 g Ag/t 35 meters 1, 2, 3
Ag 0 >45 g Ag/t 6 meters 1, 2, 3

Statistical analyses of coded assays and composites, including coefficients of variation and population-distribution plots, indicate multiple populations of significance were captured in the higher-grade domain (domain 300) of both gold and silver, as well as in the low-grade gold domain (domain 100). The recognition of multiple populations within these domains, coupled with the results of initial grade-estimation runs in which higher-grade samples in these multi-population domains were affecting inappropriate volumes in the block model, led to the use of restrictions on the search distances for the higher-grade populations of these domains. The search restrictions place limits on the maximum distances from a block that the high-grade population composites can be ‘found’ and used in the interpolation of gold and/or silver grade into that block. The final search-restriction parameters were derived from the results of multiple interpolation iterations that employed various search-restriction distances. Severe search restrictions were used for the gold and silver estimated in domain 0, as domain 0 composites of any substantive grade involve assay data that are not modeled within the mineral domains due to the lack of continuity and/or lack of geologic context.

The maximum number of composites allowed for the estimation of the low-grade domains of gold and silver are less than that of the other grade interpolations. This was done to decrease the smearing of outlier high grades that are present within these otherwise low-grade domains.

The gold and silver mineralization commonly exhibits multiple orientations throughout the DeLamar deposit. A total of 12 unique dips and 4 strike directions were identified in the DeLamar resource area. These combine to create 13 unique orientation areas distributed throughout the model area, and each area was coded into the block model using wireframed ‘estimation area’ solids.

Many of the estimation areas are characterized a single strike but two dips, which are accounted for in grade interpolation by the use of two initial passes, Pass 1 and Pass 2. The dip that reflects higher-grade mineralization was given priority, which most commonly meant the steeper of the two dips. The priority dip was then used in the search ellipse for the Pass 1 grade interpolation, while Pass 2 estimation used the secondary dip in its search ellipse. All other estimation parameters, such as search distance and sample criteria, remained identical in the two passes (Table 14.9). The third and final estimation pass was an isotropic pass, i.e., without an orientation bias, and was used to interpolate grades that were not estimated in the first two passes.

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Gold and silver grades were interpolated using inverse-distance to the third power, ordinary-krige, and nearest-neighbor methods. The mineral resources reported herein were estimated by the inverse-distance interpolation, as this method led to results that were judged to most closely respect the drill data than those obtained by ordinary kriging. The nearest-neighbor estimation was completed as a check on the inverse-distance and krige interpolations.

Grade interpolation was completed using length-weighted 3.05-meter (10-foot) composites. The estimation passes were performed independently for each of the mineral domains, so that only composites coded to a particular domain were used to estimate grade into blocks coded to that domain. Blocks coded as having partial percentages of more than one gold and/or silver domain had multiple grade interpolations, one for each domain coded into the block for each metal. The estimated grades for each gold and silver domain coded to a block were coupled with the partial percentages of those mineral domains in the block, as well as any outside, dilutionary, domain 0 grades and partial percentages, to enable the calculation of a single volume-averaged gold and a single volume-averaged silver grade for each block. These single final resource block grades, and their associated resource tonnages, are therefore fully block-diluted using this methodology.

14.8.5Model Checks

Polygonal sectional volumes derived from the sectional mineral-domain polygons were compared to the polygonal volumes derived from the level plans and long sections, as well as to the coded block-model volumes derived from the partial percentages, to assure close agreement. All block-model coding, including topographies, oxidation, estimation areas, and mineral domains, was checked visually on the computer. The nearest-neighbor and ordinary-krige estimates, as well as a polygonal grade and tonnage estimate using the cross-sectional domain polygons, were all used as a check on the inverse-distance estimation results. No unexpected relationships between the check estimates and the inverse-distance estimate were identified. The inverse-distance estimated grades were also evaluated on various grade-distribution plots that included assays, composites, and nearest-neighbor block grades as a check on both the global and local estimation results, which led to fine-tuning various estimation parameters. Finally, the inverse-distance grades were visually compared to the drill-hole assay data to assure that reasonable results were obtained.

14.9Florida Mountain Area Gold and Silver Modeling

The modeling procedures employed for the Florida Mountain resources were very similar to those used in the estimation of the DeLamar area resources (Section 14.8). The following summary of the Florida Mountain resource modeling is therefore discussed in less detail.

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14.9.1Mineral Domains

The approximate grade ranges of the low-grade (domain 100), mid-grade (domain 200), and high-grade (domain 300) grade populations and mineral domains at Florida Mountain are listed in Table 14.10.

Table 14.10 Approximate Grade Ranges of Florida Mountain Area Gold and Silver Domains

Domain g Au/t g Ag/t
100 ~0.2 to ~0.6 ~7 to ~30
200 ~0.6 to ~2.0 ~30 to ~90
300 > ~2.0 > ~90

The Florida Mountain gold and silver mineralization was modeled by interpreting gold and silver mineral-domain polygons separately on a set of vertical, 30-meter (98.4-foot) spaced, north-looking east-west cross sections that span the presently known extents of the deposit. The mineral domains were interpreted using the gold and silver drill-hole assay data, associated drill-hole lithologic codes, documented descriptions of the mineralization, Integra’s cross-sectional lithologic modeling, and wireframe solids of the historical underground workings created by MDA.

At Florida Mountain, a series of relatively thin, anastomosing, steeply dipping veins and breccias characterize the mid- to high-grade mineralization, modeled as domain 200 and 300, respectively. These thin veins and breccias are enveloped by mineralization modeled in the low-grade gold and silver domains. Taken as a whole, the mineralization forms what can be considered to be large-scale stockwork systems that are associated with a series of intrusive rhyolite ‘necks’. The continuity of any single vein or vein-breccia decreases as the grade increases, although zones characterized by these intermittent higher-grade domains do have general strike continuity, and many of these zones correlate with historically named vein zones. While the mineralization lacks continuity, especially at higher grades, the density of the drill data at Florida Mountain is sufficient to define and appropriately represent the discontinuous nature of the mineralization.

The Florida Mountain mineralization was modeled over a northly strike extent of almost 1,400 meters (4,593 feet), an east-west width of up to 675 meters (2,215 feet), and an elevation range of approximately 500 meters (1,640 feet). Cross-sections showing examples of the gold and silver mineral domains for the Florida Mountain deposit are shown in Figure 14.5 through Figure 14.8.

The final cross-sectional gold and silver mineral-domain polygons were projected three-dimensionally to the drill data in each sectional window, and these three-dimensional polygons were then sliced horizontally at eight-meter elevation intervals that match the mid-block elevations of the resource block model. The horizontal slices were used to create a new set of mineral-domain polygons for both gold and silver on level plans at eight-meter spacings that serve to rectify the domain interpretations to the drill-hole data at the scale of the block model. Level plans were used due to the steeply dipping mineralization that characterizes the entire Florida Mountain deposit.

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Figure 14.5 Florida Mountain Cross Section 2830 N Showing Geology and Gold Domains

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Figure 14.6 Florida Mountain Cross Section 2830 N Showing Geology and Silver Domains

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Figure 14.7 Florida Mountain Cross Section 3280 N Showing Geology and Gold Domains

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Figure 14.8 Florida Mountain Cross Section 3280 N Showing Geology and Silver Domains

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14.9.2Assay Coding, Capping, and Compositing

Drill-hole gold and silver assays were coded to the Florida Mountain gold and silver mineral domains using their respective cross-sectional polygons, and assay caps were defined for each domain, as well as for drill-hole assays that lie outside of the modeled domains (assigned to domain “0”), as summarized in Table 14.11. In addition to the assay caps, restrictions on the search distances of higher-grade portions of some of the domains were applied during grade interpolations (discussed further below).

Table 14.11 Florida Mountain Area Gold and Silver Assay Caps by Domain

Domain g Au/t Number Capped

(% of samples)
g Ag/t Number Capped

(% of samples)
0 5.0 32 (<1%) 100 24 (<1%)
100 3.0 10 (<1%) 70 27 (<1%)
200 9.0 6 (<1%) 175 8 (<1%)
300 75.0 13 (0.9%) 1500 9 (1.2%)

Descriptive statistics of the uncapped and capped coded assays are provided in Table 14.12 and Table 14.13 for gold and silver, respectively.

Table 14.12 Descriptive Statistics of Florida Mountain Area Coded Gold Assays

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Table 14.13 Descriptive Statistics of Florida Mountain Area Coded Silver Assays

The capped assays were composited at 3.05 meter (10-foot) down-hole intervals respecting the mineral domains. Descriptive statistics of Florida Mountain composites are shown in Table 14.14 and Table 14.15 for gold and silver, respectively.

Table 14.14 Descriptive Statistics of Florida Mountain Area Gold Composites

Table 14.15 Descriptive Statistics of Florida Mountain Area Silver Composites

14.9.3Block Model Coding

The eight-meter-spaced (26.25-foot-spaced) level-plan mineral-domain polygons were used to code a block model with a model bearing of 000° and blocks that are six meters in an east-west direction, eight meters in a north-south direction, and eight meters high. The block dimensions were increased from those used in the 2019 resource estimation and are larger than those used for the DeLamar area, reflecting conclusions derived from new geotechnical studies. The percentage volume of each mineral domain, as well as the percentage of any volume in the block lying outside the mineral domains, is stored within each block (the “partial percentages”).

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Two topographic surfaces were used to code the block model: the as-mined and present-day surfaces discussed in Section 14.2.2. These digital topographic surfaces were used to define: (1) the percentage of each block that lies within bedrock; and (2) the percentage of each block that is comprised of backfill/dump material, which lies above the as-mined surface and below the present-day surface.

The modeled mineralization has a variety of orientations, which led to the construction of wireframe solids to encompass model areas with unique orientations. These solids were then used to code the model blocks to these specific areas.

The oxidation solids described in Section 14.6 were used to code model blocks as oxide, mixed, or non-oxide. The partial percentages of the wireframe solids of the historical underground workings (see Section 14.2.3) were also coded into model blocks.

Finally, the specific-gravity values discussed in Section 14.7 were assigned to the partial percentage of bedrock coded to each model block according to the coded oxidation state of the block; the partial percentage of each block coded to backfill/dump was assigned a density value of 1.88. These coded density values were then used in combination with the stored percentages of rock and fill to determine the tonnage of each model block.

14.9.4Grade Interpolation

Multiple populations of significance were captured in the high-grade domain (domain 300) of both gold and silver, which led to the incorporation of search restrictions. Search restrictions were also used for the dilutionary material outside the mineral domains (domain 0) for both the gold and silver grade estimations.

The maximum number of composites allowed for the estimation of the low-grade domains of gold and silver in Passes 1 and 2 are less than that for all other grade interpolations. This was done to decrease the smearing of outlier grades that occur in this otherwise low-grade domain.

Gold and silver grades were interpolated using inverse distance to the third power, ordinary krige, and nearest-neighbor methods. The mineral resources reported herein were estimated by the inverse-distance interpolation, as this method led to results that were judged to more closely approximate the drill data than those obtained by ordinary kriging. The nearest-neighbor estimation was completed as a check on the inverse-distance and krige interpolations. The parameters applied to the gold-grade estimations at Florida Mountain are summarized in Table 14.16.

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Table 14.16 Summary of Florida Mountain Area Estimation Parameters

Estimation Pass – Au + Ag Domain Search Ranges (meters) Composite Constraints
Major Semi-Major Minor Min Max Max/Hole
Pass 1 – Domain 100 60 60 12 3 16 4
Pass 1 – Domain 200 + 300 + 0 60 60 12 3 25 4
Pass 2 – Domain 100 120 120 40 1 16 4
Pass 2 – Domain 200 + 300 + 0 120 120 40 1 25 4


Restrictions on Search Ranges
Domain Search Restriction Threshold Search Restriction Distance Estimation Pass
Au 300 >15 g Au/t 20 meters 1, 2
Au 0 >0.7 g Au/t 8 meters 1, 2
Ag 300 >450 g Ag/t 35 meters 1, 2
Ag 0 >30.0 g Ag/t 6 meters 1, 2

Three estimation areas were defined for the purposes of the Florida Mountain grade interpolations, with each estimation area being characterized by a unique strike orientation (350°, 345°, and 000°) and a vertical dip.

Grade interpolation was completed in two passes using length-weighted 3.05-meter (10-foot) composites. The second pass was used to estimate grades into blocks that were not estimated in Pass 1. The estimation passes were performed independently for each of the mineral domains, so that only composites coded to a particular domain were used to estimate grade into blocks coded by that domain. The estimated grades for each gold and silver domain coded to a block were coupled with the partial percentages of those mineral domains in the block, as well as the outside, dilutionary, domain 0 grades and partial percentages, to enable the calculation of a single volume-averaged gold and a single volume-averaged silver grade for each block. These single resource block grades, and their associated resource tonnages, are therefore fully block-diluted using this methodology.

14.9.5Model Checks

The model and estimation were checked in a similar manner as described for the DeLamar deposit estimation in Section 14.8.5.

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14.10DeLamar Project Mineral Resources

The DeLamar project mineral resources have been estimated to reflect potential open-pit extraction and potential processing by a variety of methods, including: crushing and heap leaching of oxide and mixed materials at DeLamar and Florida Mountain; grinding, flotation, ultra-fine regrind of concentrates, and Albion cyanide-leach processing of the reground concentrates for the non-oxide materials at DeLamar; and grinding, flotation, ultra-fine regrind of concentrates, and agitated cyanide-leaching of non-oxide materials at Florida Mountain. To meet the requirement of having reasonable prospects for eventual economic extraction by open-pit methods, pit optimizations for the DeLamar and Florida Mountain areas were run using the parameters summarized in Table 14.17 and Table 14.18 and the resulting pits were used to constrain the project resources.

Table 14.17 Pit Optimization Cost Parameters

Parameter DeLamar Florida Mtn Unit
Mining Cost $ 2.00 $ 2.00 $/tonne mined
Heap Leach
Oxide Processing $ 2.75 $ 2.75 $/tonne processed
Mixed Processing $ 3.75 $ 3.50 $/tonne processed
Incremental Haulage $ 0.20 $ 0.20 $/tonne processed
G&A $ 0.40 $ 0.40 $/tonne processed
Mill – DeLamar Area
Non-Oxide Processing $ 15.25 $ – $/tonne processed
Incremental Haulage $ 0.20 $ – $/tonne processed
G&A Cost $ 0.25 $ – $/tonne processed
Mill – Florida Mountain Area
Non-Oxide Processing $ – $ 9.00 $/tonne processed
Incremental Haulage $ 0.20 $/tonne processed
G&A Cost $ – $ 0.25 $/tonne processed
Au Price $ 1,800 $ 1,800 $/oz produced
Ag Price $ 21 $ 21 $/oz produced
Au Refining Cost $ 5.00 $ 5.00 $/oz produced
Ag Refining Cost $ 0.50 $ 0.50 $/oz produced
Royalty see Table 4.2 see Table 4.2 NSR


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Table 14.18 Pit-Optimization Metal Recoveries

DeLamar Florida Mountain
Process Type Oxide Mixed Non-Oxide Oxide Mixed Non-Oxide
Heap Leach – Au 85% 80% 90% 85%
Heap Leach – Ag 45% 40% 65% 55%
Mill – Albion – Glen Silver – Au 78%
Mill – Albion – Glen Silver – Ag 78%
Mill – Albion – Non-Glen Silver – Au 87%
Mill – Albion – Non-Glen Silver – Ag 87%
Mill – Agitated Leach – Au 95%
Mill – Agitated Leach – Ag 92%

The pit shells created using the optimization parameters were applied to constrain the project resources for both the DeLamar and Florida Mountain deposits. The in-pit resources were further constrained by the application of a gold-equivalent cutoff of 0.17 g/t to all model blocks lying within the optimized pits that are coded as oxide or mixed, a 0.3 g/t gold-equivalent cutoff for blocks coded as non-oxide at DeLamar, and a 0.2 g/t cutoff for blocks coded as non-oxide at Florida Mountain. These cutoffs were also used as overrides on the pit optimization, so that all blocks below the cutoff for each oxidation material type, if any, were treated as ‘waste’ during the optimization run.

The resource cutoff applied to non-oxide materials at Florida Mountain is lower than that at DeLamar, which is due to the lower processing costs and higher recoveries attributed to these materials at Florida Mountain.

Gold equivalency is a function of metal prices (Table 14.17) and metal recoveries (Table 14.18), with the recoveries varying by deposit and oxidation state. These variables, combined with the estimated gold and silver grades, are used to calculate a gold-equivalent grade for every block in the model. An example of the calculation of the gold-equivalent grade (“g AuEq/t”) of a Florida Mountain model block coded as mixed is as follows:

g AuEq/t = g Au/t + (g Ag/t ÷ ((1,800 x 0.85) ÷ (21 x 0.55))

where “g Au/t” and “g Ag/t” are the estimated gold and silver grades, respectively, and the other variables are the metal prices and recoveries. The gold-equivalent grades are calculated for each block for the sole purpose of applying the resource cutoffs to the appropriate materials within the optimized pits. Table 14.19 lists the gold-equivalent factors resulting from the equation above; each block’s silver grade in the model is divided by the appropriated factor in Table 14.19 and then added to the block gold grade to derive the block gold-equivalent grade to which the resource cutoff grades are applied.

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Table 14.19 Gold-Equivalency Factors Applied to Silver Grades

DeLamar Florida Mountain
Oxide Mixed Non-Oxide Oxide Mixed Non-Oxide
161 171 85 118 132 88

The total DeLamar project resources, which include the resources for both the DeLamar and Florida Mountain areas, are summarized in Table 14.20. The project mineral resources are inclusive of the mineral reserves discussed in Section 15. Mineral resources that are not mineral reserves do not have demonstrated economic viability.

Table 14.20 Total DeLamar Project Gold and Silver Resources

1.Oxide and mixed mineral resources are reported at a 0.17 g AuEq/t cutoff in consideration of potential open-pit mining and heap-leach processing. Non-oxide mineral resources are reported at a 0.3 g AuEq/t cutoff at DeLamar in consideration of potential open-pit mining and grinding, flotation, ultra-fine regrind of concentrates, and Albion cyanide-leach processing of the reground concentrates. Non-oxide mineral resources at Florida Mountain are reported at a 0.2 g AuEq/t cutoff in consideration of potential open-pit mining and grinding, flotation, ultra-fine regrind of concentrates, and agitated cyanide-leaching. These cutoffs are applied to blocks lying within optimized pits.

2.The effective date of the mineral resources is March 1, 2021.

3.Mineral resources are reported inclusive of mineral reserves.

4.Mineral resources that are not mineral reserves do not have demonstrated economic viability.

5.Rounding may result in slight discrepancies between tonnes, grade, and contained metal content.

6.The estimate of mineral resources may be materially affected by geology, environment, permitting, legal, title, taxation, sociopolitical, marketing, or other relevant issues

The current mineral resources include only the modeled mineralization that was not mined during the historical open-pit operations. The tonnage of the historical underground stopes and related workings modeled by MDA were also removed from the Florida Mountain resources.

The DeLamar project resources are classified according to the criteria presented in Table 14.21.

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Table 14.21 Resource Classification Parameters

The Measured and Indicated classification constraints for the Florida Mountain area are more restrictive than those for the DeLamar area. This is due to the differing styles of higher-grade mineralization in each of the deposits. At Florida Mountain, higher-grade mineralization occurs as irregular, large-scale stockwork zones of limited continuity that require a higher density of drilling to properly define than is the case with the more regular and continuous higher-grade mineralization at the DeLamar deposit.

Although the authors are not experts with respect to any of the following aspects of the project, the authors are not aware of any unusual environmental, permitting, legal, title, taxation, socio-economic, marketing, political, or other relevant factors not discussed in this report that could materially affect the potential development of the DeLamar project mineral resources as of the effective date of the report.

The gold and silver resources for the DeLamar and Florida Mountain areas are reported separately in Table 14.22.

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Table 14.22 Gold and Silver Resources of the DeLamar and Florida Mountain Areas

1.Oxide and mixed mineral resources are reported at a 0.17 g AuEq/t cutoff in consideration of potential open-pit mining and heap-leach processing. Non-oxide mineral resources are reported at a 0.3 g AuEq/t cutoff at DeLamar in consideration of potential open-pit mining and grinding, flotation, ultra-fine regrind of concentrates, and Albion cyanide-leach processing of the reground concentrates. Non-oxide mineral resources at Florida Mountain are reported at a 0.2 g AuEq/t cutoff in consideration of potential open-pit mining and grinding, flotation, ultra-fine regrind of concentrates, and agitated cyanide-leaching. These cutoffs are applied to blocks lying within optimized pits.

2.The effective date of the mineral resources is March 1, 2021.

3.Mineral resources are reported inclusive of mineral reserves.

4.Mineral resources that are not mineral reserves do not have demonstrated economic viability.

5.Rounding may result in slight discrepancies between tonnes, grade, and contained metal content

6.The estimate of mineral resources may be materially affected by geology, environment, permitting, legal, title, taxation, sociopolitical, marketing, or other relevant issues

Figure 14.9 through Figure 14.12 are representative cross-sections showing the estimated block-model gold and silver grades, respectively, for the DeLamar area. These figures correspond to the mineral domain cross-sections presented in Figure 14.1 through Figure 14.4.

Figure 14.13 through Figure 14.16 are representative cross-sections showing the estimated block-model gold and silver grades for the Florida Mountain area. These figures correspond to the Florida Mountain mineral domain cross-sections presented in Figure 14.5 through Figure 14.8.

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Figure 14.9 Cross Section 2190 NW Showing Sommercamp – Regan and N. DeLamar Block-Model Gold Grades

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Figure 14.10 Cross Section 2190 NW Showing Sommercamp – Regan and N. DeLamar Block-Model Silver Grades

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Figure 14.11 Cross Section 2790 NW Showing Glen Silver Block-Model Gold Grades

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Figure 14.12 Cross Section 2790 NW Showing Glen Silver Block-Model Silver Grades

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Figure 14.13 Cross Section 2830 N Showing Florida Mountain Block-Model Gold Grades

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Figure 14.14 Cross Section 2830 N Showing Florida Mountain Block-Model Silver Grades

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Figure 14.15 Cross Section 3280 N Showing Florida Mountain Block-Model Gold Grades

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Figure 14.16 Cross Section 3280 N Showing Florida Mountain Block-Model Silver Grades

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The modeled mineralization within the optimized pits that constrain the total current project resources is tabulated at various cutoffs in Table 14.23 for oxide and mixed materials, and Table 14.24 for non-oxide materials, with the current resources highlighted in bold. These tables are presented to provide grade-distribution information, which allows for detailed assessments of the project resources. The materials tabulated meet the requirement of reasonable prospects of economic extraction, as they are part of the current resources that are constrained as lying within optimized pits. As such, the mineralized materials tabulated at cutoffs higher than the resource cutoffs represent subsets of the current resources.

Table 14.23 Total Project In-Pit Oxide and Mixed Materials at Various Cutoffs

1.Rounding may cause slight discrepancies between tonnes, grade, and contained metal content.

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Table 14.24 Total Project In-Pit Non-Oxide Materials at Various Cutoffs

1.The 0.2 / 0.3 g AuEq/t cutoff tabulation reflects the resource cutoffs applied to Florida Mountain and DeLamar non-oxide materials, respectively.

2.Rounding may cause slight discrepancies between tonnes, grade, and contained metal.

14.11Discussion of Resource Modeling

The current resources described in this report differ from those reported in 2019 due to the incorporation of data from holes drilled subsequent to the 2019 resource estimations, additional constraints placed on the use of certain historical data, updated metal prices and other parameters used in the pit optimizations that constrain the resources, and resource grade cutoffs. The new drilling included 35 core holes at the DeLamar area and 54 at Florida Mountain. A significant portion of this drilling was completed in part to supply samples for metallurgical testing, and therefore the holes are representatively distributed throughout the resource areas. The new drill data resulted in slight increases in the densities assigned to the model blocks, a small increase in the size of one of the historical open pits modeled at DeLamar, minor modifications in the modeling of oxidation states, and, most substantively, increased confidence in the estimation of the current resources. Additional low-precision historical gold analyses were identified and excluded from use in the estimation of the resources. The increase in project resources from those reported in 2019 is almost entirely due to higher metal prices, higher metal recoveries at Florida Mountain, and lower resource cutoff grades for oxide and mixed materials in both deposits and non-oxidation materials at Florida Mountain.

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Consistent with assaying methods of the time period, some of the historical gold assays in the project databases were completed at a detection limit of 0.17 g Au/t (0.005 oz/ton), and the quality of historical assays that had detection limits below 0.17 g Au/t are uncertain. Current economic parameters for heap-leach processing can lead to mining cutoffs that approach or lie below this grade. Due to this uncertainty, a 0.17 g Au/t override was applied to the resource-pit optimizations, and it is also the resource cutoff grade applied to in-pit oxide and mixed blocks. Any potential future mining operation that may consider the use of mining cutoff grades below this level will need to consider this uncertainty inherent in the historical assay database.

The drilling that forms the basis of the resource estimations was done primarily by RC and conventional-rotary methods, which can be affected by down-hole contamination. As discussed elsewhere in this report, a small quantity of drill intervals in which down-hole contamination was suspected were excluded from use in the resource estimation of the DeLamar area. However, potentially contaminated samples may remain in the data used in the estimations, although the possible inclusion of such samples is not considered to be a material issue at DeLamar or Florida Mountain.

The late-1800s to early-1900s underground stopes in the DeLamar area were almost entirely mined out by the 1977 through 1998 historical open-pit mining operations. Although some of the related developmental crosscuts, etc., remain within the resources, their volumes are insignificant. At Florida Mountain, stopes and related workings along the Black Jack – Trade Dollar vein system, which were modeled by MDA, extend upwards into the Florida Mountain resource model. A total of approximately 200,000 tonnes of material lying within the modeled underground solids that would have otherwise been part of the Florida Mountain reported resources were removed from the resources.

Within the limits of the current Florida Mountain deposit resources, it is not uncommon for drill holes to have markedly different grades than adjacent holes. Mr. Gustin believes this variability is properly represented in the resource model by the explicit modeling of the gold and silver domains, combined with the tight drill spacing at Florida Mountain, where a high percentage of resource blocks lie within an average distance of 20 meters (65.62 feet) from two drill holes.

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15.0MINERAL RESERVE ESTIMATES

15.1Introduction

Mr. Dyer classifies mineral reserves in order of increasing confidence into Probable and Proven categories to be in accordance with the “CIM Definition Standards for Mineral Resources and Mineral Reserves” (2014), and therefore NI 43-101. Mineral reserves for the DeLamar and Florida Mountain deposits were developed by applying relevant economic criteria to define the economically extractable portions of the current mineral resources. CIM standards require that modifying factors be used to convert mineral resources to mineral reserves. The standards define modifying factors and Proven and Probable mineral reserves with CIM’s explanatory material shown in italics as follows:

Mineral Reserve

Mineral reserves are sub-divided in order of increasing confidence into Probable mineral reserves and Proven mineral reserves. A Probable mineral reserve has a lower level of confidence than a Proven mineral reserve.

A mineral reserve is the economically mineable part of a Measured and/or Indicated mineral resource. It includes diluting materials and allowances for losses, which may occur when the material is mined or extracted and is defined by studies at preliminary feasibility or feasibility level as appropriate that include application of modifying factors. Such studies demonstrate that, at the time of reporting, extraction could reasonably be justified.

The reference point at which mineral reserves are defined, usually the point where the ore is delivered to the processing plant, must be stated. It is important that, in all situations where the reference point is different, such as for a saleable product, a clarifying statement is included to ensure that the reader is fully informed as to what is being reported.

The public disclosure of a mineral reserve must be demonstrated by a preliminary feasibility study or feasibility study.

Mineral Reserves are those parts of Mineral Resources which, after the application of all mining factors, result in an estimated tonnage and grade which, in the opinion of the Qualified Person(s) making the estimates, is the basis of an economically viable project after taking account of all relevant Modifying Factors. Mineral Reserves are inclusive of diluting material that will be mined in conjunction with the Mineral Reserves and delivered to the treatment plant or equivalent facility. The term ‘Mineral Reserve’ need not necessarily signify that extraction facilities are in place or operative or that all governmental approvals have been received. It does signify that there are reasonable expectations of such approvals.

‘Reference point’ refers to the mining or process point at which the Qualified Person prepares a Mineral Reserve. For example, most metal deposits disclose mineral reserves with a “mill feed” reference point. In these cases, mineral reserves are reported as mined ore delivered to the plant and do not include reductions attributed to anticipated plant losses. In contrast, coal reserves have traditionally been reported as tonnes of “clean coal”. In this coal example, mineral reserves are reported as a “saleable product” reference point and include reductions for plant yield (recovery). The Qualified Person must clearly state the ‘reference point’ used in the Mineral Reserve estimate.

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Probable Mineral Reserve

A Probable mineral reserve is the economically mineable part of an Indicated mineral resources, and in some circumstances, a Measured mineral resource. The confidence in the modifying factors applying to a Probable mineral reserve is lower than that applying to a Proven mineral reserve.

The Qualified Person(s) may elect, to convert Measured Mineral Resources to Probable Mineral Reserves if the confidence in the Modifying Factors is lower than that applied to a Proven Mineral Reserve. Probable Mineral Reserve estimates must be demonstrated to be economic, at the time of reporting, by at least a preliminary feasibility study.

Proven Mineral Reserve

A Proven mineral reserve is the economically mineable part of a Measured mineral resource. A Proven mineral reserve implies a high degree of confidence in the modifying factors.

Application of the Proven mineral reserve category implies that the Qualified Person has the highest degree of confidence in the estimate with the consequent expectation in the minds of the readers of the report. The term should be restricted to that part of the deposit where production planning is taking place and for which any variation in the estimate would not significantly affect the potential economic viability of the deposit. Proven mineral reserve estimates must be demonstrated to be economic, at the time of reporting, by at least a preliminary feasibility study. Within the CIM Definition standards the term Proved Mineral Reserve is an equivalent term to a Proven Mineral Reserve.

Modifying Factors

Modifying Factors are considerations used to convert mineral resources to mineral reserves. These include, but are not restricted to mining, processing, metallurgical, infrastructure, economic, marketing, legal, environmental, social and governmental factors.

Mr. Dyer has used Measured and Indicated mineral resources as the basis to define mineral reserves for both the DeLamar and Florida Mountain deposits. Mineral reserve definition was done by first identifying ultimate pit limits using economic parameters and pit optimization techniques. The resulting optimized pit shells were then used for guidance in pit design to allow access for equipment and personnel. Mr. Dyer then considered mining, processing, metallurgical, infrastructure, economic, marketing, legal, environmental, social, and governmental factors for defining the estimated mineral reserves.

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The DeLamar mine plan has been developed using five pit phases. The Florida Mountain mining has been designed using three pit phases. Mr. Dyer used the phased pit designs for both deposits to define the project production schedule, which was then used for cash-flow analysis for the PFS. The final cash-flow model demonstrates that the deposits make a positive cash flow and are reasonable with respect to statement of mineral reserves for those deposits.

15.2Pit Optimization Parameters

Pit optimizations were completed by first identifying economic and geometric parameters. Pit optimizations were then completed using Lerchs-Grossman techniques to create nested sets of pit shells, and economic analysis was done within the pit shells.

15.2.1Economic Parameters

Economic parameters were used to generate optimized pits using a Lerchs-Grossman algorithm within Whittle™ software (Version 4.7). The economic parameters include mining costs, processing costs, general and administrative costs (“G&A”), refining costs, royalties, and metal recoveries. Mine planning is an iterative process, and initial costs and recoveries were assumed to determine how large pits would be. The economic parameters were refined as concepts were developed on how material would be processed from the two separate deposits. The methods for processing that were determined include:

  • Use of crushing and cyanide heap leaching for oxide and mixed material from DeLamar and Florida Mountain; and

  • Using milling for higher-grade, non-oxide material from DeLamar and Florida Mountain.

The economic parameters used are shown in Table 15.1. The overall heap-leaching process rate is planned to be 35,000 tonnes per day or 12,600,000 tonnes per year for both Florida Mountain and DeLamar oxide and mixed material. DeLamar heap-leach processing will also include agglomeration. Initially only the oxide and mixed material will be processed from both deposits, then starting in year 3, non-oxide will be processed primarily from the Florida Mountain deposit and then from the DeLamar deposit through a plant constructed to operate at a rate of 6,000 tonnes per day or 2,160,000 tonnes per year. The G&A assumes a cost of $6.0 million per year and was weighted-averaged to assume $0.44 per tonne of heap-leach and $0.22 per tonne of non-oxide material. In the final cash-flow model the G&A cost was applied on a per year basis.

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Table 15.1 DeLamar and Florida Mountain Economic Parameters

GMV = gross metal value; COG = cutoff grade.

Royalties were applied by royalty area or region as provided by Integra. These are described in Section 4.3.

Recoveries were applied based on recommendations by Mr. Jack McPartland, as summarized in Section 13.0. Recoveries are shown in Table 15.2. The oxide and mixed recoveries assume crushed heap leaching for oxide and mixed material, and flotation milling for non-oxide material. Florida Mountain non-oxide material uses recovery Equation 1 and Equation 2 to estimate the recoveries based on gold and silver grades respectively.

Table 15.2 DeLamar and Florida Mountain Recoveries

Equation 1 Florida Mountain Gold Recovery

Where: Maximum recovery = 87%

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Equation 2 Florida Mountain Silver Recovery

Where: Maximum recovery = 77%

15.2.2Cutoff Grades

Cutoff grades were applied based on a block value calculated using metal price and recoveries by area and oxidation. The cost to process the material was then used to calculate a cutoff grade. The metal prices used to calculate the block value were $1,650 per ounce of gold and $21.00 per ounce of silver, though the final cash-flow model used $1,700 per ounce gold and $21.50 silver. The block value calculation equation is shown in Equation 3. The block value was calculated for both gold and silver and then each was added together to provide an overall value for each block.

Equation 3 Block Value Calculation

The recoveries that were used are shown in Table 15.2, Equation 1, and Equation 2.

The cutoff grades used are internal cutoff grades meaning that they do not include the cost of mining. The pit designs were based on economic pit shells, which include the cost of mining. Since all the material inside of the pit is to be mined, the mining cost becomes a sunk cost and the point where the decision is made to process the block is at the top of the ramp.

Cutoff grades applied represent the cost to process material along with G&A and incremental haulage costs and are shown along with the economic parameters used in Table 15.1. Note that royalties are built into the block values as shown in Equation 3, and are considered in determining whether to process the material. In addition, the DeLamar non-oxide breakeven cutoff grade would be $11.44/t but a cutoff of $15.00 was used as an elevated cutoff grade to enhance the project’s economic performance.

15.2.3Geotechnical Parameters

RESPEC conducted a geotechnical investigation and engineering analysis to develop recommendations for bench-face angles (“BFAs”), inter-ramp angles (“IRAs”), overall slope angles (“OSA”s), and catch-bench widths for the recommended planned pit slopes. This investigation and engineering analysis was conducted under the supervision of Mr. Jay Nopola of RESPEC’s Rapid City, South Dakota office. The results were reported in Raffaldi et al. (2021) and were based on the following information:

  • Historical geotechnical data;

  • Laboratory rock-mechanics tests;

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  • Visual observations of the performance on existing highwalls;

  • Verbal and written communications with current and former site employees;

  • Estimates of rock mass properties from available rock core and core photographs;

  • Geological maps and three-dimensional geological models;

  • High-resolution unmanned aerial vehicle (“UAV”) point-cloud data; and

  • The 2019 Preliminary Economic Assessment (“PEA”) of Gustin et al. (2019) and associated pit designs.

Note that the PEA pit designs were used as reference for the geotechnical studies as a general guide to the depth and extents of the PFS pits to guide the geotechnical studies. The resulting recommendations were completed to a PFS level of study and were applied to the PFS of this technical report. Reference to the PEA pits below is made to reflect the extents and depth of the PFS study. It is anticipated that additional geotechnical studies will be completed in the future.

Historical geotechnical data included laboratory rock mechanics test results, rock-mass properties, and discontinuity orientations. The efforts undertaken in the mid- to late-1980s were considerable and the methodologies employed are well-described in historical reports and are largely in line with current standards of practice. Additional laboratory strength testing and geotechnical site investigation conducted by RESPEC were in good agreement with the historical data allowing it to be incorporated in the PFS level evaluation and providing confidence in the geotechnical model which forms the basis for the PFS recommendations.

DeLamar Main, Sullivan Gulch, Milestone, and Florida Mountain pit areas were subdivided into sectors based on geology and expected pit slope height. Kinematic and limit-equilibrium slope stability analyses were conducted, and recommendations were given by pit sector and material type. The geology expected in the pit areas and pit sectors are shown in Figure 15.1 and Figure 15.2. The Milestone pit is not shown but was treated as a single geotechnical sector. Milestone is located to the northwest of the DeLamar deposit.

RESPEC’s PFS-level recommendations for pit slope design at DeLamar and Florida Mountain are summarized in Table 15.3 and Table 15.4. Recommended layback angles for slopes in soil materials throughout the site are summarized in Table 15.5.

Joint sets in the DeLamar and Florida Mountain areas generally dip steeply (>70 degrees) or nearly horizontally (<20 degrees). As expected, joints in both areas are also usually oriented subparallel or orthogonal to regional fault trends. These conditions allowed for steeper BFA recommendations than used for the PEA-level design. However, presplit or other controlled-blasting methods and careful scaling and excavation practices will be required to ensure design conformance and clean bench faces.

Primary factors impacting achievable overall slope angles from Raffaldi et al. (2021) include:

(1)The location and width of fault zones, and groundwater elevations and pore water pressures;

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(2)The location and extent of weak, highly altered rock. Zones of weak rock are commonly associated with alteration classifications of moderate and strong argillization (arg2, arg3). The distribution of these materials is currently understood only in terms of broad generalizations. Particularly for the DeLamar area, the project geological model should be updated so that the spatial distribution of alteration can be viewed in geological cross sections and used directly in geotechnical evaluations.

(3)The location and width of fault zones. The location and geotechnical character of faults should be better characterized. Major faults that are expected to intersect the final pit slopes should be evaluated using the best-available interpretation of orientations, fault zone widths, and geotechnical characteristics. All fault data should be plotted and its influence on slope stability evaluated considering intersections with potentially problematic contacts and weak rock units.

(4)Groundwater elevations and pore water pressures. Groundwater conditions are currently based on limited historical information. A hydrogeologic study should be performed to measure and/or verify expected groundwater conditions in each of the PEA pits. To support the hydrogeologic study and help characterize seasonal fluctuations in pore water pressures, sealed piezometers should be installed in a select number of future exploratory boreholes. The piezometers should be near the final highwalls (upslope) and central sectors of the proposed pits. The hydrogeological consultant should use the hydrogeological model developed for pit dewatering predictions to assess the distributions of pore pressures behind the pit slopes at critical locations and times and develop pit dewatering plans. Slope stability should be re-evaluated with this data.

Many of these suggestions are underway as of the effective date of this report. Future geotechnical data collection will focus on refining these aspects of the geotechnical model and may allow for steeper overall slope angles.

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Figure 15.1 DeLamar Main- Sullivan PFS Geotechnical Sectors on 2019 PEA Pits

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Figure 15.2 Florida Mountain PFS Geotechnical Sectors on 2019 PEA Pit

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Table 15.3PFS Pit Slope Design Recommendations – DeLamar Main, Sullivan Gulch, and Milestone

Sector Inter-Ramp

Slope Angle


(deg)
Average

Bench-Face

Angle


(deg)
Final Bench

Height


(m)
Design Catch

Bench Width


(m)
Northwest 40 70 12 9.9
Southwest 40 70 12 9.9
North Central 45 70 12 7.6
North and South Wahl 40 70 12 9.9
Sommercamp and North DeLamar 45 70 12 7.6
South DeLamar 40 70 12 9.9
Sullivan Gulch 45 70 12 7.6
Milestone 45 70 12 7.6

Table 15.4 PFS Pit Slope Design Recommendations – Florida Mountain

Sector Inter-Ramp

Slope Angle


(deg)
Average

Bench-Face

Angle


(deg)
Final Bench

Height


(m)
Design Catch

Bench Width


(m)
East (Tuff) 30 70 16 22.0
East (Tql/Tpr) 45 70 16 10.2
West 50 70 16 7.7
South 50 70 16 7.7


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Table 15.5 PFS Slope Design Recommendations for Soil Materials

Material

(Maximum Height)
Maximum Overall

Slope Angle

(Horizontal:Vertical)
Qcl (≤ 60 m) 1.75:1
Qcl (≤ 90 m) 2:1
Tal, Hbx (≤ 90 m) 2:1
Qbf (≤ 90 m) 2:1

15.2.4Royalty Boundaries

Royalty zone ownership and boundaries were provided by Integra as summarized in Table 15.6 and Figure 15.3 for Florida Mountain and Figure 15.4 for DeLamar. The royalty rates are summarized in Section 4.3.

Table 15.6 Royalty Zones Ownership

The boundaries shown in Figure 15.3 and Figure 15.4 were flagged into the resource block model along with the prevailing rates. These are also carried into the production schedule and financial model.

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Figure 15.3 Florida Mountain Royalty Zones

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Figure 15.4 DeLamar Royalty Zones

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15.3Pit Optimization with Pit by Pit Analysis

Pit optimizations were run using Whittle™ software (version 4.7). Inputs into Whittle included the resource block model along with the economic, geotechnical, and royalty parameters previously discussed. Each deposit was run separately, and ultimate pit shells were selected from the Whittle results for final designs. For DeLamar and Florida Mountain, additional pit shells were considered for guidance of interior pit phases.

The selections of ultimate pits and pit phases were done as a two-step process. The first step was to optimize a set of pit shells based on varying a revenue factor. This was done in Whittle using a Lerchs-Grossman (“LG”) algorithm. The revenue factor was multiplied by the recovered ounces and the metal prices, essentially creating a nested set of pit shells based on different metal prices. Revenue factors for each of the deposits were varied from 0.30 to 2.0 in increments of 0.025 with a base price of $1,000 per ounce of gold, so the resulting pit shells represent gold prices from $300 to $2,400 per ounce in increments of $25.00. This has the potential of generating up to 89 different pit shells that can be used for analysis.

The second step of the process was to use the Pit by Pit (“PbP”) analysis tool in Whittle to generate a discounted operating cash flow (note that capital is not included). This uses a rough scheduling by pit phase for each pit shell to generate the discounted value for the pit. The program develops three different discounted values: best, worst, and specified. The best-case value uses each of the pit shells as pit phases or pushbacks. For example, when evaluating pit 20, there would be 19 pushbacks mined prior to pit 20, and the resulting schedule takes advantage of mining more valuable material up front to improve the discounted value. Evaluating pit 21 would have 20 pushbacks; pit 22 would have 21 pushbacks and so on. Note that this is not a realistic case as the incremental pushbacks would not have enough mining width between them to be able to mine appropriately, but this does help to define the maximum potential discounted operating cash flow.

The worst case does not use any pushbacks in determining the discounted value for each of the pit shells. Thus, each pit shell is evaluated as if mining a single pit from top to bottom. This does not provide the advantage of mining more valuable material first, so it generally provides a lower discounted value than that of the best case.

The specified case allows the user to specify pit shells to be used as pushbacks and then schedules the pushbacks and calculates the discounted cash flow. This is more realistic than the base case as it allows for more mining width, though the final pit design will have to ensure that appropriate mining width is available. The specified case has been used for each of the two DeLamar project mines to determine the ultimate pit limits to design to, as well as to specify guidelines for designing pit phases.

15.3.1DeLamar Pit Optimization

The previously discussed parameters were used along with base metal prices of $1,650 per ounce of gold and $21.00 per ounce of silver. Gold prices were varied from $300 to $2,500 per ounce to create the pit optimization results for the DeLamar deposit. The silver price was adjusted to maintain a constant gold to silver price ratio of $1,650/oz Au to $21.00/oz Ag. The pit optimization results are shown in Table 15.7 in $100 gold price increments with the addition of a $1,650 pit results. The $1,650 and $1,700 pit shells are highlighted in the table. The $1,650 price was used to determine the ultimate pit limits for design. The $1,700 price was used for the PFS economic evaluation as discussed in Section 22, thus this pit shell shows potential upside value.

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Table 15.8 shows the PbP results for pit shells 29 through 61. The highlighted pit is the one that maximizes the discounted operating cash flow using a $1,650 gold price and $21.00 silver price and is used as the basis for the DeLamar ultimate pit design. Figure 15.5 shows the results of the DeLamar PbP analysis graphically.

Table 15.7 DeLamar Pit Optimization Results

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Table 15.8 DeLamar Pit by Pit Results

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Figure 15.5 DeLamar Pit by Pit Graph

15.3.2Florida Mountain Pit Optimization

Florida Mountain pit optimizations also used the parameters previously discussed and the optimizations were completed with gold prices varying from $300 to $2,500 per ounce to create 89 pit shells. These results are shown in Table 15.9 in $100 gold price increments with the addition of a $1,650 and $1,700 pit shell, which are highlighted in the table.

Table 15.10 shows the PbP results for pit shells 29 through 61. The highlighted pit is the one that maximizes the discounted operating cash flow and is used as the basis for the pit designs. Figure 15.6 shows the results of the Florida Mountain PbP analysis graphically.

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Table 15.9 Florida Mountain Pit Optimization Results

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Table 15.10 Florida Mountain Pit by Pit Results

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Figure 15.6 Florida Mountain Pit by Pit Graph

15.4Road and Ramp Design

Road designs have been completed for the PFS to allow primary access for people, equipment, and consumables to the site. This includes haul roads between the designed pits, dumps, and proposed leach and mill facilities. Within the pit designs, ramps have been established for haul truck and equipment access. The in-pit ramps will only require a single berm. Ramps outside of the pit will require two safety berms. One-lane traffic ramps are also utilized near the bottom of pits where the strip ratio is minimal, and the traffic requirements are low.

The ramps and haul roads assume the use of Komatsu 136-tonne haul trucks with an operating width of 7.0 meters (23 feet). Haul roads were designed using 34 meters (112 feet) for two-way traffic and were narrowed to 25 meters (82 feet) for one-way traffic in deeper portions of the pits where the strip ratio per bench is minimal.

Road designs are intended to have a maximum of 10% gradient, though some may exceed this for short distances around inside turns. Where switchbacks are utilized, the centerline gradient is reduced to about 8%. This keeps the inside gradient approximately 12%. Switchback designs do not have added detail for super elevation through the curves, but is it assumed that this will be done when they are constructed.

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15.5Pit Design

Pit designs were completed for DeLamar and Florida Mountain using Surpac™ software (version 6.7). Slope parameters utilize recommendations from the Raffaldi et al. (2021) geotechnical study. The recommendations are summarized in Section 15.2.3.

Some slopes will be through soil-like materials for limited heights. These slopes will be designed using between a 1.75 meter to 2.0 meter (5.7-foot to 6.6-foot) horizontal distance for each meter of height.

Florida Mountain designs were created using 8-meter (26.2 feet) benches. Catch bench width ranging from 7.7 meters (25.3 feet) in the west and south of the pit and 10.2 meters (33.5 feet) in the east except for tuffs where a 22-meter (72.2-feett) catch bench was used. The bench face angle used is 70°. The resulting inner-ramp slopes range from 30° in the tuff, to 45° slopes in the east, and 50° slopes in the south and west.

Florida Mountain pit designs were completed using three pit phases. Phase one is mined as a smaller pit in the northeast. Phase two is mined around phase one on the eastern side and then the final phase is mined to the west. These designs are shown in Figure 15.5, Figure 15.6, and Figure 15.7. The Florida Mountain pit designs contain a mixture of heap-leach and mill ore with most (81% heap-leach ore and 19% mill ore). Proven and Probable reserves by pit phase are summarized in Section 1.1.

DeLamar pit designs utilize 6-meter (19.7-feet) benches with a catch bench width ranging from 7.6 meters (24.9) to 9.9 meters (32.5 feet) installed every other bench, or 12 meters (39.4 feet) within rock slopes. The bench face angle used is 70°. The resulting inner-ramp to inter-ramp slopes range from 40° to 45°.

DeLamar pit designs utilize five pit phases to establish a mining sequence. The first mining at DeLamar would start in the Milestone pit where the waste material will be used as construction material for the tailing facility. The next mining would be done in the DeLamar main pit phase 1, followed by phase 2, and then phase 1 and phase 2 from the Sullivan Gulch pit. The DeLamar pit designs are shown in Figure 15.10 through Figure 15.14. The DeLamar Proven and Probable reserves by pit phase are summarized in Section 1.7 and Section 15.6.

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Figure 15.7 Florida Mountain Phase 1 Design

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Figure 15.8 Florida Mountain Phase 2 Design

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Figure 15.9 Florida Mountain Phase 3 and Ultimate Pit Design

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Figure 15.10 DeLamar – Milestone Phase 1 Pit Design

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Figure 15.11 DeLamar Main Pit, Phase 1 Design

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Figure 15.12 DeLamar Main Phase 2 and Ultimate Pit Design

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Figure 15.13 DeLamar – Sullivan Gulch Phase 1 Pit Design

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Figure 15.14 DeLamar – Sullivan Gulch Phase 2 and Ultimate Pit Design

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15.6Proven and Probable Reserves

Mr. Dyer estimated mineral reserves for each deposit based on the pit designs shown in Figure 15.7 through Figure 15.14. Measured resources were converted to Proven reserves and Indicated resources were converted to Probable reserves. Total Proven and Probable reserves for the DeLamar project from all pit phases are 123,483,000 tonnes at an average grade of 0.45 g Au/t and 23.27 g Ag/t, for 1,787,000 ounces of gold and 92,403,000 ounces of silver (Table 15.11). The mineral reserves point of reference is the point where is material is fed into the crusher .

Florida Mountain Proven and Probable reserves are shown by pit phase in Table 15.12 and the Florida Mountain Proven and Probable reserves are summarized by oxidation and category types in Table 15.14.

DeLamar Proven and Probable reserves are summarized by pit phase in Table 15.13 and the DeLamar Proven and Probable reserves are summarized by oxidation and category types in Table 15.15.

Proven and probable reserves are also shown by deposit and process type in Table 15.16.

Table 15.11 Total Proven and Probable Reserves, DeLamar and Florida Mountain

Notes:

(1)All estimates of mineral reserves have been prepared in accordance with National Instrument 43 – 101 – Standards of Disclosure for Mineral Projects (“NI 43-101”) and are included within the current Measured and Indicated mineral resources.

(2)Thomas L. Dyer, PE for MDA, a division of RESPEC, in Reno, Nevada, is a Qualified Person as defined in NI 43-101, and is responsible for reporting Proven and Probable mineral reserves for the DeLamar Project. Mr. Dyer is independent of Integra.

(3)Mineral reserves are based on prices of $1,650 per ounce Au and $21.00 per ounce Ag. The reserves were defined based on pit designs that were created to follow optimized pit shells created in Whittle. Pit designs followed pit slope recommendations provided by RESPEC.

(4)Reserves are reported using block value cutoff grades representing the cost of processing (See Equation 3):

Florida Mountain oxide leach cutoff grade value of $3.55/t.

Florida Mountain mixed leach cutoff grade value of $4.20/t.

Florida Mountain non-oxide mill cutoff grade value of $10.35/t.

DeLamar oxide leach cutoff grade value of $3.65/t

DeLamar mixed leach cutoff grade value of $4.65/t.

DeLamar non-oxide mill cutoff grade value of $15.00/t.

(5)The mineral reserves point of reference is the point where is material is fed into the crusher.

(6)The effective date of the mineral reserves estimate is January 24, 2022.

(7)All ounces reported herein represent troy ounces, “g Au/t” represents grams per gold tonne and “g Ag/t” represents grams per silver tonne.

(8)Columns may not sum due to rounding.

(9)The estimate of Mineral reserves may be materially affected by geology, environment, permitting, legal, title, taxation, sociopolitical, marketing, or other relevant issues.

(10)Energy prices of US$2.50 per gallon of diesel and $0.065 per kWh were used.

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Table 15.12 Florida Mountain Reserves by Pit Phase

Table 15.13 DeLamar Reserves by Pit Phase

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Table 15.14 Florida Mountain Proven and Probable Reserves

Table 15.15 DeLamar Proven and Probable Reserves

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Table 15.16 Proven and Probable Reserves by Process Type

Mr. Dyer is not aware of any mining, metallurgical, infrastructure, permitting, and other relevant factors that could materially affect the estimated mineral reserves beyond those discussed in other sections of this report. Mr. Dyer is not aware of any legal, political, environmental, or other risks that could materially affect the potential development of the mineral reserves.

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16.0MINING METHODS

The PFS presented in this report considers open-pit mining of the DeLamar and Florida Mountain gold-silver deposits. Mining would utilize 23-cubic meter (30-cubic yard) hydraulic shovels along with 13-cubic meter (16.7-cubic yard) loaders to load 136-tonne capacity haul trucks. The haul trucks would haul waste and ore out of the pit and to dumping locations. Due to the length of ore hauls, the ore will be stockpiled near the pits followed by loading into a Railveyor system which will convey the ore into a crusher. The Railveyor system will be supplemented with haul trucks on an as needed basis.

Waste material will be stored in waste-rock storage facilities (“WRSFs”) located near each of the Florida Mountain and DeLamar deposits, as well as backfilled into pits where available. The exception is the Milestone pit, from which waste material will be fully utilized for construction material for the tailing storage facility (“TSF”).

This section describes the locations and designs for WRSFs and backfill designs, mine production schedule, process material delivery schedule, stockpiling schedule, mine equipment requirements, and personnel requirements for the PFS.

16.1WRSFs and Backfill Designs

WRSFs, along with backfill areas, have been designed for the PFS to contain the waste material mined from the different pit phases. Volumes of waste material mined from the Florida Mountain and DeLamar pits are shown in Table 16.1. The volumes of waste material are calculated using the SG value in the resource model along with a 1.3 swell factor. The swell factor is intended to represent the swelling of material as it is blasted and loaded into trucks along with recompacting the material as it is placed in the WRSFs and backfill areas. The WRSFs are shown in the GA drawing in Section 18.1.

Table 16.1 Waste Rock Containment Requirements (With Swell)

Note: FlMtn = Florida Mountain; Del = DeLamar.

Swell factor of 1.3 was used for containment requirements

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Waste rock storage capacities are shown in Table 16.2. A single WRSF design is planned for Florida Mountain along with a two backfill dumps into the Florida Mountain phase 1 and 2 pits. Material from Florida Mountain phase 1 will be placed into the primary WRSF. Phase 2 waste material will also be placed into the primary WRSF except for some upper areas of the pit where some waste will be backfilled. Phase 3 waste material is planned to be placed into the backfill dump as available while the remaining waste material will be placed into the Florida Mountain WRSF. The total capacity of the WRSF is 32.2 million cubic meters (42.1 million cubic yards). The remaining 23.4 million cubic meters (30.6 million cubic yards) of waste material will be placed into backfill.

Three WRSF designs were created for the DeLamar area which includes a West WRSF, East WRSF, and a North WRSF. The West and East WRSFs are intended for storage of material from the DeLamar Main phase 1 pit. Both dump designs include a roadway that will be built into the WRSFs to allow haulage through the main pit exits for both DeLamar Main and Sullivan Gulch pits. The East WRSF creates its haulage road through a valley to the south of the deeper Sullivan Gulch phase 2 pit. This road is anticipated to be in place well before the mining of Sullivan Gulch phase 2. The total West DeLamar WRSF total capacity is 5.9 million cubic meters (7.7 million cubic yards). After the roadway is completed, the East WRSF is to be expanded to the south. The total East DeLamar WRSF total capacity will be 50.0 million cubic meters (65.4 million cubic yards).

The North WRSF will be located in a valley to the north of the Main and Sullivan Gulch pits. This will be used for the Main pit phase 2 waste along with Sullivan Gulch pit waste. The designed capacity of the North WRSF is 26.4 million cubic meters (34.5 million cubic yards). As available, additional waste will be placed into the Main phase 1 pit and from the Main phase 2 pit as backfill. Additional backfill material will be placed into the Main phase 2 pit from Sullivan Gulch phase 1 mining.

Table 16.2 WRSF and Backfill Design Capacities

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16.2Mine Production Schedule

Production scheduling was completed using Geovia’s MineSched™ (version 2021) software. Proven and Probable reserves along with waste material inside pit designs previously discussed were used to schedule mine production.

The production schedule considers the processing of Florida Mountain oxide and mixed material by crushing and heap leaching. Florida Mountain non-oxide material would be processed using flotation followed by cyanide leaching of the flotation concentrate. Processing of the DeLamar material will require crushing and agglomeration prior to heap leaching and non-oxide material will be processed through the mill.

Monthly periods were used to create the production schedule with pre-stripping starting in Florida Mountain at month -5. The start of leach processing is scheduled in month 1, though a total of 610,000 tonnes of leach material is to be mined during preproduction. It is assumed that this material will be crushed and used as over liner on top of the leach pad liner. The nominal rate for leach processing will be 35,000 tonnes per day or 12,600,000 tonnes per year. Note that during the first year, a total of 10,287,000 tonnes will be processed along with the material laid onto the pad during preproduction. This represents a ramp up to full processing.

The leach pad stream of material will not always maximize the process. Full throughput is dependent on availability of leach ore.

Leaching starts with Florida Mountain material in month 1 and DeLamar leach material is processed starting in year 2. Prior to that, the agglomeration circuit will be installed. The DeLamar leach material will be processed up to the same rate as the Florida Mountain material.

Florida Mountain and DeLamar non-oxide material will be stockpiled until the flotation mill is constructed. The start of the 6,000 tonne per day mill will be in year 3 with 1,982,000 tonnes processed in that year, increasing to 2,160,000 tonnes per year after that until the non-oxide material is exhausted.

The total mining rate would ramp up from an initial 2,000 tonnes per day to about 60,000 tonnes per day over a period of six months. A maximum of 138,000 tonnes per day is used in later years when the stripping requirement becomes more significant in Florida Mountain phase 3.

The yearly mining production for Florida Mountain and DeLamar is summarized in Table 16.3 and Table 16.4, respectively. Table 16.5 summarizes the total yearly mine production schedule.

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Table 16.3 Florida Mountain Mine Production Schedule

COG = cutoff grade.

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Table 16.4 DeLamar Mine Production Schedule

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Table 16.5 Total PFS Mine Production Schedule

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The material sent to the crusher for heap-leach processing was scheduled by MDA based on the mine production schedule and metal recoveries. The estimated heap-leach recovered ounces were modeled by Integra’s metallurgical process consultant, Mr. Michael Botz, as reported in Section 17.0. The recoveries used to estimate recoverable gold are those shown in Table 15.2. Leach K Ozs Au Rec and K Ozs Ag Rec in Table 16.6 shows recoverable ounces based on the recoveries provided. The yearly process production summary in Table 16.6 also shows the tonnage, grade, contained ounces of silver and gold, and the recovered silver and gold ounces from mill processing.

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Table 16.6 PFS Process Production Schedule

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Stockpiles will be located near the Railveyor loadout facilities. The stockpiles will be used to store low-grade material longer term, as well as some higher-grade material during initial mining. Stockpile management will be required not only to be able to manage the metal grades to be processed but will also be critical to manage blending of the various ore types. Stockpile management will require additional studies in the future to ensure optimization of the mine.

Table 16.7 and Table 16.8 show stockpile balance sheets for the heap-leach and mill material, respectively.

Table 16.7 Leach Ore Stockpile Balance

Table 16.8 Mill Ore Stockpile Balance

16.3Equipment Requirements

The PFS has assumed owner mining instead of the more expensive contract mining. The production schedule was used along with additional efficiency factors, performance curves, and productivity rates to develop the first-principal hours required for primary mining equipment to achieve the production schedule. Primary mining equipment includes drills, loaders, hydraulic shovels, and haul trucks.

Support, blasting, and mine maintenance equipment would be required in addition to the primary mining equipment. Table 16.9 shows the yearly equipment requirements.

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Table 16.9 PFS Yearly Mine Equipment Requirements

Note that earlier mining studies identified the water table to be around the 1,810 meter (5,938 foot) elevation. All of the Florida Mountain mining and most the DeLamar mining is above this elevation. The mining at Sullivan Gulch phase 2 does extend about 160 meters (525 feet) below the 1,810 elevation. It is assumed that the two pit pumps will be sufficient to maintain a dry pit. Additional studies will be completed as part of a feasibility study.

The mine is anticipated to operate 24 hours per day, utilizing four crews of workers each working four days on and four days off. It is anticipated that these crews would rotate between day shift and night shift. The daily shift schedule would be 12 hours per day reduced to account for standby time including startup/shutdown, lunch, breaks, and operational delays totaling 3.0 hours per day. This allows for 21 work hours in each day or 87.5% schedule efficiency. The estimated schedule efficiency is shown in Table 16.10.

Table 16.10 Schedule Efficiency

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Pioneer drills would be smaller air-track drills with contained cabs and the production drills are anticipated to be 45,000lb-pulldown, track-mounted, rotary blast-hole drills. An 83% efficiency factor was used for pioneer drilling and 85% efficiency was used for production and controlled blast-hole drilling. Penetration rates of 26.3, 27.5, 30.3 meters per hour (86.3, 90.2, 99.4 feet per hour) were used along with 2.8, 2.8, and 3.0 minutes per hole of non-drilling times for production, trim-rows, and pioneer drilling, respectively.

Based on the parameters used, one pioneer drill and five production drills are estimated to be needed. It is assumed that these drills will last through the life of mine (“LOM”) with an availability of 85% for the life of the drill.

Loading equipment is anticipated to include one large 13-cubic meter (17 cubic yard) loader and two 23-cubic meter (30 cubic yard) hydraulic shovels. The loader theoretical productivity was estimated to be 2,345 tonnes per hour, or 1,950 tonnes per hour at an operating efficiency of 83%. The loader is primarily used for back-up mining production and re-handle of material from stockpiles. The assumed availability starts at 90% and is reduced 1% per year until it reaches 85%, and then is held constant through the life of the loading units. No replacement loaders were assumed. The overall use of available hours is 33%.

Two hydraulic shovels are used as the primary loading tool. The initial shovel starts operating in month -6 and the second shovel starts working in month 1. The theoretical productivity was estimated to be 3,326 tonnes per hour, or 2,760 tonnes per hour after applying 83% efficiency. As with the loader, the assumed availability starts at 90% and declines at 1% per year to a low of 85%, and then remains the same through the LOM. The overall use of operating hours is 80%.

Haul trucks are assumed to be 136-tonne capacity rigid frame trucks. Haulage hours were developed using MineSched software (Version 2021). MineSched uses 3-dimensional centerlines drawn for bench, in-pit, and ex-pit travel. The performance and retard curve data are input into the software, and MineSched uses that along with the truck capacity and load, dump, and spot times to determine the time required to haul material to its destination. The hours developed from MineSched are considered productive hours, and these are adjusted in the mining cost spreadsheets to include an 83% efficiency.

The loading time provided in the software is based on the hydraulic shovel and are included in the productive hour calculation. This is adjusted in spreadsheets to reflect the use of loaders; thus the load time is dependent on whether the truck was loaded by a loader or shovel. The loader time used was 3.73 minutes and the shovel time used was 2.70 minutes. Spot time at the loader or shovel was 0.50 minutes and the spot and dump time was a combined 1.20 minutes. A capacity of 131 tonnes per load was used as dry tonnage to reflect the dry densities in the resource block model. The number of trucks was calculated to increase over time due to farther haulage with some pit phases. A total of 16 haul trucks are purchased to maintain the production schedule. This assumes a 1% per year declining availability from 90% down to 85%.

Railveyor will primarily be used for haulage of ore material from stockpiles near the pits to the crusher feeding both the heap-leach pad and the mill. Discussion on Railveyor along with the layout is provided in Section 18.6.

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16.4Personnel Requirements

Table 16.11 shows the estimated mine operations personnel requirements (full mine site personnel requirements are shown in Table 18.3). This is based on the number of people that will be required to operate, supervise, maintain, and plan for operations to achieve the production schedule. The peak mining personnel requirement is 250 people on an annual basis.

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Table 16.11 PFS Mining Personnel Requirements

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17.0RECOVERY METHODS

This section was prepared by Mr. Benjamin Bermudez of M3 Engineering in Tucson, Arizona, and Mr. Art Ibrado of Fort Lowell Consulting in Tucson, Arizona. Some US customary units are used in this section because equipment and design components are customarily specified in US customary units rather than metric units.

The proposed process facilities for the DeLamar project comprise a heap-leach operation to process oxide and mixed ores, and a mill to process non-oxide ore. The heap-leach operation will have a nominal capacity of 35,000 tonnes per day while the mill will have a nominal capacity of 6,000 tonnes per day.

The heap-leach operation will employ a conventional heap-leach process, which will include three-stage crushing, agglomeration, and stacking on multi-lift dedicated leach pads. The mill process will include primary crushing, semi-autogenous (“SAG”) milling, ball milling, flotation, very fine grinding of flotation concentrates, cyanide leaching of the concentrates, and counter-current decantation (“CCD”). Pregnant leach liquors from both heap-leach and mill operations will be processed by common Merrill-Crowe and refinery facilities, to precipitate gold and silver from solution and smelt the precipitate to produce doré bullions.

17.1Process Production Schedule

The process facilities for the DeLamar project will be developed to accommodate the mining sequences of the Florida Mountain and DeLamar deposits as summarized in Section 16. A preliminary mine schedule is shown in Table 17.1. The LOM average head grades to the heap-leach operations are estimated to be 0.40g Au/t and 17.3g Ag/t. The expected LOM average grades for the milling operations are 0.61g Au/t and 40.91g Ag/t.

The mine schedule indicates that the heap leach will operate throughout the mine life, receiving ore initially from Florida Mountain, lasting though Year 7. DeLamar ore deliveries will start in Year 2 and continue through Year 17. The mill will start receiving non-oxide ore from both Florida Mountain and DeLamar in Year 3 and will continue until Year 17.

17.2Process Design

The flowsheets developed for the DeLamar project PFS are based on the metallurgical testing and interpretation presented in Section 13. The following subsections provide a summary of the main components of the process design criteria, a description of the PFS process flowsheets, the major process equipment selected for the project, the primary buildings required to support the major process equipment, a description of the primary process support infrastructure including the water systems, power, process air systems, and the tailing handling system.

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The primary crushing system mass flow rates are based on 75% plant availability and the combined required capacity to maintain both oxide and non-oxide stockpile levels. The oxide secondary and tertiary crushing system mass flow rates are based on 80% plant availability through ore stacking on the pad. The non-oxide ore mass flow rates are based on a plant availability of 92%. For simplicity, the estimated availability used in sizing each unit operation is a combination of mechanical availability and equipment utilization and, therefore, reflect estimated operating times. Every operation may have its own definitions that may differ according to its needs.

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Table 17.1 Heap Leach and Mill Feed Schedules

Year Heap-leach Ore Mill Ore
Florida Mountain DeLamar Total Florida Mountain DeLamar Total
kt g/t Au g/t Ag kt g/t Au g/t Ag kt g/t Au g/t Ag kt g/t Au g/t Ag kt g/t Au g/t Ag kt g/t Au g/t Ag
PreProd 610 0.42 10.55 610 0.42 10.55
1 10,287 0.50 13.43 10,287 0.50 13.43
2 10,596 0.46 10.44 1,953 0.44 20.85 12,548 0.46 12.06
3 9,474 0.37 8.16 3,161 0.45 30.08 12,635 0.39 13.64 1078 1.09 15.57 904 0.54 77.47 1,982 0.84 43.80
4 10,993 0.42 11.76 79 0.30 12.10 11,072 0.42 11.77 1885 0.77 19.35 275 0.47 58.17 2,160 0.73 24.30
5 3,386 0.39 16.20 9,214 0.35 12.71 12,600 0.36 13.65 1863 0.67 19.56 297 0.62 53.40 2,160 0.67 24.21
6 15 0.33 24.54 12,585 0.38 18.50 12,600 0.38 18.51 926 0.38 8.00 1,234 0.89 28.74 2,160 0.67 19.85
7 348 0.16 11.22 10,566 0.32 21.91 10,914 0.31 21.57 558 0.28 6.74 1,608 0.54 47.40 2,166 0.48 36.94
8 6,325 0.29 33.16 6,325 0.29 33.16 445 0.28 6.74 1,715 0.46 54.19 2,160 0.43 44.41
9 997 0.19 44.62 997 0.19 44.62 492 0.28 6.74 1,668 0.52 47.50 2,160 0.47 38.21
10 296 0.46 24.02 296 0.46 24.02 740 0.28 6.74 1,420 0.43 42.39 2,160 0.38 30.18
11 450 0.60 57.84 450 0.60 57.84 346 0.28 6.74 1,819 0.60 63.44 2,166 0.55 54.37
12 259 0.70 77.92 259 0.70 77.92 379 0.28 6.74 1,781 0.74 61.76 2,160 0.66 52.11
13 176 0.79 77.57 176 0.79 77.57 300 0.28 6.74 1,860 0.78 67.77 2,160 0.72 59.28
14 170 0.74 78.07 170 0.74 78.07 307 0.28 6.74 1,853 0.86 66.58 2,160 0.78 58.07
15 200 0.61 74.82 200 0.61 74.82 277 0.28 6.74 1,889 0.78 60.26 2,166 0.71 53.41
16 103 0.48 71.64 103 0.48 71.64 302 0.28 6.74 1,858 0.68 55.00 2,160 0.62 48.25
17 0.008 0.78 143.5 0.008 0.78 143.5 656 0.28 6.74 506 0.43 23.45 1,162 0.35 14.01
Total or Average 45,708 0.43 11.40 46,533 0.36 23.10 92,241 0.40 17.30 10,555 0.53 12.26 20,688 0.65 55.52 31,243 0.61 40.91


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17.3Heap Leach Operation

The proposed initial processing facility is designed to process the Florida Mountain and DeLamar oxide and mixed ore types using conventional heap leaching for the extraction of precious metals and solution recovery using Merrill-Crowe zinc cementation process. Table 17.2 lists the major design criteria for the heap-leach process. The proposed process is depicted in the simplified flow sheet shown in Figure 17.1. The arrangement of the heap-leach facilities is shown in Figure 17.2.

Table 17.2 Heap Leach Major Design Criteria

Parameter Florida Mountain DeLamar
Processing Scheme Crush/Heap Leach Crush/Agglomerate/Heap Leach
Crushing Circuit Configuration Three Stage Crushing Three Stage Crushing with Drum Agglomeration
Crush Size 80% -12.7 mm 80% -12.7 mm
Heap Stacking Method Overland Conveyor with Radial Stacker Overland Conveyor with Radial Stacker
Leach Cycle, days 120 120
Lift Height, m 10 10
Solution Application Rate, L/hr/m2 6.1 6.1
Nominal Barren Solution Flow, m3/hr 2,119 2,119
Solution Application Method Drip Drip
Precious Metal Recovery Method Merrill-Crowe Zinc Precipitation Merrill-Crowe Zinc Precipitation
Nominal PLS Flow Rate to Plant, m3/hr 2,040 2,040

17.3.1Heap-Leach Crushing Plant and Agglomeration

The DeLamar heap-leach crushing plant will comprise three stages of crushing, starting with a gyratory crusher for primary crushing, followed by a standard cone crusher for secondary crushing, and finally, two short-head cone crushers for tertiary crushing. Table 17.3 is a list of the major equipment in the crushing plant.

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Figure 17.1 Simplified Flow Sheet of the DeLamar Project Heap-Leach Facility

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Figure 17.2 General Layout of the DeLamar Project Process Facilities

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Table 17.3 List of Main Mechanical Equipment for the Heap-Leach Crushing Plant

Equipment Number Description Installed kW
Primary Crusher 1 Gyratory Crusher, 42″ x 65″; 152 mm (6-inch) OSS* 448
Secondary Crusher 1 Standard Head Cone Crusher; Metso MP1250 (o/e)*; CSS* = 35 mm (1-3/8-inch) 933
Tertiary Crusher 2 Short Head Cone Crusher; Metso MP1250 (o/e); CSS = 19 mm (3/4-inch) 933 (each)
Secondary Screen 1 Inclined, Vibrating, Double Deck; 1,823 dMTPH (Flow Sheet); 3 m x 7.3 m) (10 ft x 24 ft) (est); Apertures: 76.2 mm (3 inches), 31.75 mm (1.25-inch) 112
Tertiary Screens 2 Inclined, Vibrating, Single Deck; 2,349 dMTPH (Flow Sheet); 3.66 m x 7.31 m (12 ft x 24ft) (est); Apertures: 21 mm (0.82-inch) 112 (each)
Primary Crusher Discharge Conveyor 1 2,278 dMTPH, 152 cm (60-inch) wide, 72.2 m (237 ft) long, 5.2 m (17 ft) lift 224
Coarse-Ore Stockpile Feed Conveyor 1 2,278 dMTPH, 152 cm (60-inch) wide, 285 m (934 ft) long, 24.4 m (80 ft) lift 746
Secondary Screen Feed Conveyor 1 1,823 dMTPH, 122 cm (48-inch) wide, 191.4 m (628 ft) long, 15.2 m (50 ft) lift 373
Transfer Conveyors 2 4,698 dMTPH, 122 cm (48-inch) wide, 192 m (630 ft) long, 3 m (10 ft lift; and 169 m (555 ft) long, 24.4 m(80 ft) lift 298 & 933
Crushing Circuit Product Transfer Conveyor 1 1,823 dMTPH, 122 cm (48-inch) wide, 247 m (810 ft) long, 6 m (20 ft) lift 261
Agglomeration Feed Conveyor 1 Batch Use, 1,823 dMTPH, 122 cm (48-inch) wide 75
Agglomerated Product Conveyor 1 Batch Use, 1,823 dMTPH, 122 cm (48-inch) wide 75
Agglomeration Drum 1 1,823 dMTPH; 3.66 m Dia x 12.2 m Length, Tire Driven; w/ VFD 261
*OSS = open-side setting; CSS = closed-side setting; o/e = or equivalent

A specialized ROM conveyance system (“Railveyor”) will primarily deliver ROM to the primary crusher dump pocket. The crushed product will drop to the primary crusher product surge bin, which is equipped with an apron feeder that will deliver the ore to the primary discharge conveyor and on to a diverter cart. The diverter cart will allow primary crushed oxide and mixed ore to be routed to the oxide coarse-ore stockpile via a dedicated oxide ore stockpile feed conveyor, or non-oxide ore to be routed to the non-oxide ore stockpile via a separate, non-oxide ore stockpile feed conveyor.

Three apron feeders, two operating and one standby, will reclaim ore from the heap-leach coarse-ore stockpile and load it to the secondary screen feed conveyor. The secondary screen will separate +32 millimeter (1.25-inch) material and feed it to the secondary crusher. The screen undersize drops to a transfer conveyor and joins the secondary crusher product to the tertiary crusher feed bin.

From the tertiary crusher feed bin, the material is split between two tertiary screens with an aperture of 21 millimeters (0.82-inch). The oversize is fed to the tertiary crushers, whose products are recycled back to the tertiary crusher feed bin and on to the tertiary screens. The undersize of the tertiary screen is the final product of the crushing plant and has an estimated size of 80% finer than 12.7 millimeters (0.5-inch).

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The crushing plant product is discharged to the crushing circuit product transfer conveyor where lime is added from a lime silo. The ore is then transferred to the stacking system feed conveyor or, alternatively, to the agglomeration feed conveyor, via a diverter cart.

Ore agglomeration will be achieved mainly at the transfer points of the grasshopper conveying system. Moisture will be supplied to the process by adding fresh water to the ore on the stacking system feed conveyor. For about 45% of the DeLamar oxide ore, agglomeration will be performed in an agglomeration drum using raw water or barren solution and cement, which will be added to the ore on the crushing circuit product transfer conveyor.

17.3.2Stacking and Heap Leaching

The final crusher or agglomerator product reports to the stacking system feed conveyor that feeds the first of the grasshopper conveyors. The initial stacking system will comprise 20 grasshopper conveyors, and an index feed conveyor and a horizontal index feed conveyor that feeds the heap-leach stacker conveyor. The initial conveying and stacking equipment are listed in Table 17.4.

Table 17.4 List of Conveying and Stacking Equipment for the Heap-Leach Pad

Equipment Number Description Drive, kW
Stacking System Feed Conveyor 1 1,823 dMTPH 298
Grasshopper Conveyors 20 1,823 dMTPH 112
Index Feed Conveyor 1 1,823 dMTPH 149
Horizontal Index Conveyor 1 1,823 dMTPH 112 (each)
Heap Leach Stacker Conveyor 1 1,823 dMTPH 298

The heap-leach pad will be a dedicated pad, which will be built in stages as mining progresses. The heap-leaching facility will consist of two dedicated heap-leach pads, namely the Jacob’s Ridge leach pad and the Valley leach pad. Each leach pad will have its own pregnant leach solution (“PLS”) pond. The PLS pond at the Jacob’s Ridge leach pad will be sized to contain the operating volume, the drain down volume, and a 100-year 24-hour precipitation event, plus freeboard.

After stacking, pipe headers and drip irrigation lines will be added to the heap surface. Sodium cyanide solution will be applied to the heap surface via the header/drip system at a proposed application rate of 6.10 L/hr/m2 for the preliminary leach cycle of 120 days. The cyanide solution, applied at a nominal flow rate of 2,119 cubic meters per hour (9,329.7 gpm) to the heap surface, will percolate though the heap until it reaches the impervious leach pad liner at the bottom of the heap. The PLS will flow by gravity to the collection point of the leach pad and on to the pregnant solution pond for each leach pad. From the pregnant ponds, the PLS will be pumped to the clarifier filter feed tank in the Merrill-Crowe plant. The heap-leach solution application rate may be adjusted during the leach cycle depending on the available area for irrigation.

Heap-leach modeling performed by Integra’s metallurgical consultant, Mr. Michael Botz of Elbow Creek Engineering Inc. in Billings, Montana indicates the upper-end sustained flow rates of barren solution and pregnant solution will be approximately 1,950 and 1,810 cubic meters per hour (8,585.6 gpm to 7,969.2 gpm), respectively. These flow rates are predicted for the period when the heaps are stacked with ore at about 35,000 t/d. The difference of 140 cubic meters per hour (616 gpm) between the flows arises from evaporation plus uptake of solution by the ore for initial wetting. The design of the Merrill-Crowe metal recovery circuit must accommodate the predicted pregnant solution flow rate of 1,810 cubic meters per hour (7,969.2 gpm). With an added 10% design margin, the hydraulic sizing basis for the Merrill-Crowe circuit would be approximately 2,000 cubic meters per hour (8,805.7gpm).

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For the case where a tank leaching circuit is operated at the site along with the heap leach, an additional 100 cubic meters per hour (440 gpm) of pregnant solution would also be routed to the Merrill-Crowe circuit. This would give a total predicted feed flow rate to the Merrill-Crowe circuit of 1,910 cubic meters per hour (8,409.5 gpm). With an added 10% design margin, the hydraulic sizing basis for the Merrill-Crowe circuit would be 2,100 cubic meters per hour (9,250gpm). M3 Engineering has selected a flow rate of 2,119 cubic meters per hour (9,330 gpm) for purposes of this PFS.

The design of the leach pad and solution ponds are discussed in more detail in Section 18.0.

17.3.3Heap-Leach Solution Pond Operation

There are three solution ponds and all will be constructed in Phase 1. One of the solution ponds will be located at the north end of the Jacob’s Ridge heap and two ponds will be located at the north (lower) end of the valley heap. The pond at Jacobs Ridge is sized to contain an operating volume of 15,140 cubic meters (4,000,000 gallons) of pregnant solution storage, plus 100% of a 100-year, 24-hour precipitation event on the ridge portion of the heap, a volume of 24,000 cubic meters (6,342,400 gallons), plus a draindown volume of 24,224 cubic meters (6,400,000 gallons), plus the volume of 0.6 meters (2.0 feet) of freeboard, 7,066 cubic meters (1,867,000 gallons). As dimensioned, the overall volume of the Jacobs Ridge pond is 71,642 cubic meters (18,928,000 gallons). This pond is designed to never overflow. Following a large storm event, the pond should be pumped at a minimum of 568 liters per second (9,000 gpm) and if there is excess pumping capacity over that needed for application to the heap, the extra flow should be routed to the concentrate leach TSF for temporary storage.

The two ponds at the lower (north) end of the valley are an operating pond and an event pond. The operating pond will be double lined with leak detection and will contain 44,920 cubic meters (11,868,000 gallons) of pregnant solution storage as an operating volume, plus 0.6 meters of freeboard. Excess flows from a precipitation event up to a 100-year, 24-hour event will flow through a spillway from the operating pond to a single-lined event pond where they will be temporarily stored until they can be pumped into the operating pond as makeup pregnant solution for return to the plant. The capacity of the event pond will be 71,130 cubic meters (18,792,659 gallons).

Pregnant solutions will be pumped to the process area in a welded steel pipeline following the access road from the valley pond area, past the Jacobs Ridge pond area to the plant for extraction of precious metals.

17.3.4Heap-Leach Production Forecasting

Gold and silver production forecasts for the Jacob’s Ridge and Valley heap-leach pads were provided by Elbow Creek Engineering. The forecasts were developed using a heap-leach production model that follows the modeling technique described by Botz and Marsden (2019). The dynamic model provides monthly estimates of gold and silver productions according to the planned construction and operation of the heaps. The forecasts take into account kinetic leach reactions, transport of leached metal to the heap liner, and holdup of leached metal within the heap pore moisture inventory. Kinetic leach rates were sourced from column testwork performed by McClelland for each ore type. Ultimate (final) extractions for gold and silver by ore type were also provided by McClelland. Scale-up discount factors were applied to the extractions by McClelland to account for full-scale leach inefficiencies, plus the small residual of leached metal that will be retained in the heap pore moisture and the end of operations.

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A summary of the gold and silver forecasts for the heap-leach pads is provided in Table 17.5. The total forecast LOM gold recovery is estimated to be 843,800 oz, and the total forecast LOM silver recovery is estimated to be 19,168,000 oz.

With any heap-leach production forecast model, there will always be a variance between modeled and actual metal productions. This variance arises from inherent variabilities associated with many aspects of heap operation, including variability in ultimate extractions, leach times, lift heights, irrigation practices, mineralogy, particle size distributions, permeability, etc. The variances that could generally be expected for a well-understood heap are shown in Table 17.6 (Marsden and Botz 2017).

Table 17.5 Heap-leach Gold and Silver Production Forecasts

Year Gold [oz] Silver [oz]
Year Cumulative Year Cumulative
1 9,900 9,900 152,000 152,000
2 119,400 129,300 1,956,000 2,108,000
3 138,700 268,000 1,863,000 3,971,000
4 111,600 379,600 1,715,000 5,686,000
5 107,700 487,300 1,867,000 7,553,000
6 103,200 590,500 1,838,000 9,391,000
7 101,300 691,800 2,184,000 11,575,000
8 78,300 770,100 2,606,000 14,181,000
9 42,500 812,600 2,336,000 16,517,000
10 10,400 823,000 1,285,000 17,802,000
11 2,700 825,700 105,000 17,907,000
12 5,800 831,500 333,000 18,240,000
13 3,300 834,800 254,000 18,494,000
14 2,900 837,700 187,000 18,681,000
15 2,500 840,200 171,000 18,852,000
16 2,300 842,500 188,000 19,040,000
17 1,200 843,700 110,000 19,150,000
18 100 843,800 11,000 19,161,000
19 0 843,800 4,000 19,165,000
20 0 843,800 3,000 19,168,000
Total 843,800 19,168,000


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Table 17.6 Typical Achievable Variances Between Modeled and Actual Metal Productions

Forecast Period 1σ Level 2σ Level
Monthly ±20% ±40%
Quarterly ±15% ±30%
Annually ±6% ±12%
Life of Heap ±1% ±2%

17.4Milling Operations

The design of the mill for non-oxide ore includes a primary crusher, which is shared with the oxide ore crushing plant, a grinding circuit, flotation, very fine grinding of concentrate, and a cyanidation leach/CCD circuit for the recovery of precious metals. Table 17.7 is a list of the main design criteria for the mill. The proposed simplified process flow sheet for the mill is shown in Figure 17.3. A list of major equipment in the mill is given in Table 17.8.

17.4.1Comminution

Primary crushing of non-oxide ore will be performed in the same gyratory crusher as the heap-leach operation. Coarse ore will be retrieved and sent to a separate non-oxide coarse ore stockpile, from which ore is reclaimed by two reclaim feeders, and transferred to the SAG mill feed conveyor.

The grinding circuit comprises a SAG mill and a ball mill. The SAG mill will operate in a closed circuit with a vibrating screen. The oversize of the screen will be conveyed back to the SAG mill via the screen oversize conveyors and SAG mill feed conveyor. The undersize slurry will flow into the ball mill cyclone feed pump box as fresh feed to the ball mill circuit.

The ball mill will operate in a closed circuit with the primary hydrocyclone cluster, which sends the underflow back to the ball mill and the overflow to the flotation circuit via the trash screen. The target grind for the ball mill circuit is 80% finer than 150 microns.

Possible future additions to the circuit include a pebble crusher and flash flotation.

17.4.2Flotation and Regrind

Ground ore will first be conditioned with potassium amyl xanthate (“PAX”), ethyl sec-butyl dithiophosphate (“Aerofloat 208”) and frother (“MIBC”) in an agitated tank for 15 minutes at 30% solids. Sodium metasilicate may also be added as a dispersant with high-clay ores. Once conditioned, the slurry flows by gravity to the rougher flotation (four cells) and rougher scavenger flotation (three cells) banks. Additional flotation reagents will be introduced to the slurry between the rougher and rougher scavenger banks. The combined rougher and rougher scavenger residence time is 45 minutes, to produce a combined flotation concentrate at an estimated mass pull of 10% for DeLamar ores and 5% for Florida Mountain ores.

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Other reagents are added to the flotation process depending on the ore being processed. These are included in Table 17.9

Future plans for the flotation plant could include two stages of cleaning to further upgrade the concentrate prior to regrinding.

Table 17.7 Main Design Criteria for Concentrator and Cyanidation Plants

Parameter Value
Milling – Non-Oxide Material Type
Overall Milling Scheme Comminution/Flotation/Conc. Leach
Comminution Circuit Configuration Primary Crusher, SAG Mill and Ball Mill
Crush Size 80% -120 mm
Primary Grind Size 80% -150 µm
Flotation Cell Type Tank Cell
Flotation Circuit Configuration Rougher-Rougher Scavenger
Rougher Flotation Retention Time, min 25
Rougher Scavenger Retention Time, min 20
Flotation Tailing Handling Method Thicken, Tailing Storage Facility
Concentrate Mass (for regrind) Percent of Total Feed 10%
Concentrate Leach Solids Density, % wt:wt 40
Concentrate Leach Retention Time 24 hrs.
Regrind Size 80% -20 µm
Au Recovery
Florida Mountain 83%
DeLamar 37%
Ag Recovery
Florida Mountain 72%
DeLamar 75%
Leach Tailing Handling Method Concentrate Leach Tailing Storage Facility


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Figure 17.3 Simplified Process Flow Diagram of the Concentrator and Cyanide Leach Plants

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Table 17.8 List of Main Mill Equipment

Equipment Number Description Drive, kW
Primary Crusher 1 Uses primary crusher at Oxide Ore Processing 448
SAG Mill Feed Conveyor 1 340 dMTPH; 169 m (553 ft) L, 4.3 m (14 ft) Drop 112
Semi-Autogenous (SAG) Mill 1 6.7 m diam x 4.57 m EGL* (22 ft x 15 ft), with internal grates & pulp discharge 2,984
SAG Mill Discharge Screen 1 Inclined, Vibrating, Single Deck; 340 dMTPH (Flow Sheet); 3 m x 6 m (10 ft x 20 ft) (est); Apertures: 8 mm 75
Ball Mill 1 4.57 m diam x 9.14 m EGL (15 ft x 30 ft) 2,984
Primary Hydrocyclone Cluster 1 1,697 m3/h (7,472 gpm); 10-place Cluster (est); 20-inch Cyclones; 5-Operating; 3-Stand-by; 2-Spare Ports n/a
Rougher Flotation 4 100 m3 Tank Cell (est) 112 (each)
Rougher Scavenger Flotation 3 100 m3 Tank Cell (est) 112 (each)
Utlra-fine Regrind Mill 1 ISAMill Model M5000 (o/e) 1,119
Regrind Cyclone Cluster 1 91.4 m3/h (402 gpm); 10-place Cluster (est); 6-inch Cyclones; 8-Installed, 2 Spare Ports n/a
Pre-Leach Thickener 1 12.2 m (40 ft) Diameter; High Rate 15
Leach Tanks 7 218 gpm; 24-hr Circuit Retention; 20ft D x 24ft H; 54k gal Total per Tank; Open Top; rubber-lined carbon steel 19 (each)
Post-Leach Thickener 1 12.2 m (40 ft) Diameter; High Rate 15
Pregnant Solution Clarifier 1 12.2 m (40 ft) Diameter; High Rate 15
CCD Cyclone Clusters 4 107 – 158 m3/h (470 – 697 gpm); 10-place Cluster (est); 6-inch Cyclones; 8-Installed, 2 Spare Ports n/a
CCD Mix Tanks 4 158 m3/h (697 gpm); 5-min Retention; 3 m dia x 3.66 m H (10 ft x 12 ft); Open Top; rubber-lined carbon steel 3.7 (each)
Cyanide Destruction Tank 2 42 m3/h (185 gpm); 60-min Retention; 3.66 m D x 4.88 m H (12 ft x 16 ft; Open Top; rubber-lined carbon steel 19 (each)

EGL is effective grinding length.

The rougher scavenger flotation tailing will flow by gravity to the flotation thickener to be dewatered to 55% solids by weight. The thickener overflow is to be pumped to the process water tank. The thickened underflow is to be pumped to the tailing storage facility. A portion of the rougher flotation tailing will be blended with concentrate leach tailing at a 1:1 solid mass ratio to improve the settling characteristics of the finer tailing.

The combined flotation concentrate will proceed to the regrind circuit, which consists of an ISAMill and a regrind cyclone cluster. The concentrate will enter the circuit through the regrind cyclone feed tank and will then be pumped to the regrind cyclone cluster. The underflow of the regrind cyclones will be fed by gravity to the ISAMill, which operates in open circuit. The overflow of the regrind cyclones and the product of the ISAMill both report to the preleach thickener, which will thicken the leach feed to 50% solids. The target grind of the leach feed is 80% finer than 20 microns. The pre-leach thickener overflow will be pumped to the process water tank.

17.4.3Concentrate Leaching

The concentrate leaching circuit comprises six leach tanks operating in series followed by a CCD system. The leach tanks are designed to provide a residence time of four hours each, for a total leach time of 24 hours. Milk-of-lime (“MOL”) is added to slurry at the pre-leach thickener feed while sodium cyanide is added to the leach feed splitter box with stage addition of either into the leach tanks as required. Air is introduced below the agitators of each leach tank, supplied by one leach air blower.

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The final leach residue will be pumped to the CCD circuit, which starts with a post-leach thickener and a series of four CCD cyclone clusters, which are supported by four agitated CCD mix tanks and four pumps (one operating per cluster). The cyclone clusters in series serve the same purpose as thickeners in series, with the post-leach thickener serving in the position of the first CCD stage. Overflow from the post-leach thickener is clarified in a thickener clarifier before being pumped to the Merrill-Crowe plant.

The underflow of the final CCD cyclone cluster will be pumped to the CCD cyanide destruction plant to destroy cyanide with metabisulfite and air. Once detoxified, the concentrate leach tailing is deposited in a separate concentrate leach tailing storage facility. In the future, concentrate leach tailing may be mixed with the fresh heap-leach feed in an agglomeration drum to extend its leach time and reduce the need for tailing detoxification and separate tailing storage facility.

17.5Merrill-Crowe Plant and Refinery

The PLS reporting to the pregnant solution pond and, starting year 3, solution from the CCD circuit pregnant solution clarifier will be pumped to the clarifier filter feed tank at the Merrill-Crowe plant. Solution clarification will be performed by three clarifying filters arranged to operate in parallel, with two operating and one on standby. The filters will have been precoated with diatomaceous earth (“DE”) to aid filtration. The solution may be infused with DE as body feed when required. The clarified solution then proceeds to the deaeration tower where it will be introduced into an evacuated chamber to remove as much dissolved oxygen as possible. After deaeration, powdered zinc, cyanide, and lead nitrate will be added to the solution to initiate an exchange redox reaction where zinc metal loses electrons to gold and silver, thereby reducing gold and silver to their metallic state and oxidizing zinc to form cyano complexes in solution.

The mixture will then be pumped to three recessed plate and frame filters operating in parallel, two operating and one standby. Precipitation of gold and silver by the exchange reaction continues as the solution makes its way to the Merrill-Crowe precipitate filters and reaches completion inside the filters. All the precipitated gold and silver will remain in the filter presses until they are discharge when the filters are full. The filtrate solutions, stripped of values, will report to the barren solution tank. Additional cyanide and caustic may be introduced into the barren solution tank before it is recycled to the heap and to the leaching tanks and CCD circuit when the mill is operating.

Gold and silver precipitates collected by the filter presses will be dried in a retort to remove moisture and mercury before they are fluxed and smelted in an induction melting furnace. At the end of smelting, molten metal is poured into bullion molds to produce the final plant product, doré bars, which are packed for shipment.

The slag, which is poured first into a slag pot, will be weighed, sampled and stored for further processing if required, or transferred to the heap or fed to the ball mill when the mill is operating.

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Off gases from the retort furnace will be treated to remove mercury by condensation. The remaining gas is subsequently passed through a bed of sulfur-impregnated carbon. Off gases from the induction furnace will be filtered, scrubbed with water sprays and finally passed through a bed of sulfur-impregnated carbon.

A settling pond will receive backwash water from the clarifier filters to settle diatomaceous shale for solution recycling. It will also receive water from the refinery wet scrubber.

17.6Reagents

The main reagents to be used in leach processes in the heap-leach pad and the mill will be lime and sodium cyanide. In addition, the mill will use a suite of flotation reagents consisting of two flotation collectors, a frother, and two other conditioning reagents as required. The Merrill-Crowe process will use reagents that are standard in typical Merrill-Crowe operations. Table 17.9 is a summary of the reagents projected to be used in this project including points of addition and estimated consumption rates.

The MOL system consists of a lime slaker and MOL distribution tank. Pebble lime will be stored in a silo and metered into a vertical grinding mill by a screw feeder. The MOL produced will be pumped to a 4.3-meter (14-foot) diameter by 4.9-meter (16-foot) high MOL storage/distribution tank. MOL will be pumped to the solution regeneration tanks through a MOL loop.

Pebble lime will also be added to the oxide and mixed ores from another lime silo to the crusher product transfer conveyor.

Table 17.9 Main Process Reagents and Consumables

Reagent Area or Point of Addition Dosage Dosage unit
Sodium Cyanide Leach, Barren Solution Pond & Merrill-Crowe 0.3 to 0.6 kg/t
Mill – Cyanidation 0.26 to 1.25 kg/t
Caustic Soda Sodium Cyanide Distribution Tank 0.005 to 0.025 kg/t
Lime Heap Leach 0.3 to 2.4 kg/t
Mill 0.1 to 0.7 kg/t
Cyanide Destruction 0.6 to 1.2 kg/t conc
Cement Heap Leap – Agglomeration 2.8* kg/t
Zn Dust, estimate Merrill-Crowe 25.8 kg/kOz Au
47.1 kg/kOz Ag
Lead Nitrate, estimate Merrill-Crowe 15 ppm in PLS
Diatomaceous Earth (DE), estimate Merrill-Crowe 45.4 kg/filter batch
Melting Flux, estimate Refining 5.5 g/oz of metal
Flocculant, estimate Tailing Thickener, Pre-Leach Thickener, & Post-Leach Thickener 20 to 30 g/t solids
Antiscalant, estimate Process Water Pump, Barren Solution Pump 5 g/t of solution
Potassium Amyl Xanthate (PAX) Flotation 0.025 kg/t
Ethyl sec-butyl dithiophosphate (Aerofloat 208) Flotation 0.05 Kg/t


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Reagent Area or Point of Addition Dosage Dosage unit
Methyl Isobutyl Carbinol (MIBC) Frother) Flotation 0.1 kg/t
Sodium Carbonate Flotation 0.25** kg/t
Sodium Silicate (for high-clay ores) Flotation 0.025*** kg/t
Sodium Metabisulfite Cyanide Destruction 1 to 2 kg/t conc
Copper Sulfate Flotation/Cyanide Destruction 0.016 kg/t conc

*Assumes 45% of DeLamar ore requires agglomeration.

**Factored (assuming required for 25% of ore)

***Factored (assuming required for 10% of ore)

17.7Tailing Storage Facilities (TSF) and Water Consumption

As most operating mines, as much water as possible is planned to be recycled from slurry thickeners and mill tailing. Water lost to evaporation, heap-leach ore, and mill tailing needs to be replenished by raw water. In addition, raw water is used for dust suppression on mine roads and potentially for potable water.

17.7.1Tailing Storage Facilities and Water Reclamation

The plans for tailing storage for the mill operations include two facilities – one to store flotation tailing, and the other, a smaller facility, to store concentrate leach tailing.

Water reclaimed from the flotation TSF will report to the process water tank, to combine with the tailing thickener overflow, pre-leach thickener overflow, and make-up raw water coming from the oxide circuit freshwater system. The process water tank will supply water to the grinding circuit, rougher flotation, flocculant make-down, and regrind circuits.

The smaller tailing storage facility will be dedicated to concentrate leached tailing to isolate the stream within the cyanidation loop. Water reclaimed from this TSF will be pumped to the CCD circuit to supply part of the wash water for CCD.

17.8Water Consumption

The DeLamar project is projected to require a total average of 249.6 cubic meters per hour (1,099 gpm) of raw water makeup to sustain the operation, for a total yearly requirement of 2.2 million cubic meters (581 million gallons). The water usage is broken down as follows:

  • Heap-leach operation103.7 cubic meters per hour (457 gpm);

  • Milling operation130.2 cubic meters per hour (573 gpm); and

  • Mine dust suppression15.7 cubic meters per hour (69 gpm).

The actual requirement will fluctuate according to the season. A discussion of the site-wide water balance is presented in Section 18.8.

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17.9Blowers and Compressors

The design of the leach pad and mill facilities includes blowers and compressors to provide general plant air, aeration for flotation, leach, detoxification, dust control, and instrument air. A list of air equipment is given in Table 17.10.

Table 17.10 List of Blowers and Compressors for Supply Plant and Instrument Air

Equipment Tag Service Area Power, kW
200-CM-003 Air Compressor for Fine Crushing and Dust Collection 56
500-DC-014-BL Refinery Wet Scrubber Blower 75
900-CM-004 Merrill-Crowe Plant Air Compressor 56
320-CM-003 Grinding Area Compressed Plant Air Supply and Dust Collection 56
420-BL-003 Rougher Flotation Blower 261
420-BL-005, -006 Cleaner Flotation Blower (Future); 1 operating, 1 standby 149 (ea)
420-CM-004 Cleaner Flotation Area Compressed Plant Air (Future) 37
440-BL-001 Leach Air Blower, 1360 Nm3/h (800 scfm), 138 kPa (20 psig) 75
460-BL-009 CCD Cyanide Destruction Blower, 1020 Nm3/h (600 scfm), 97 kPa (14 psig) 30
820-BL-005 Lime Unloading Blower, batch use 11

17.10Power Consumption

The total connected power load is estimated at 20,907 kW for the heap-leach operation and 12,473 kW for the mill. The connected power in each process area is given in Table 17.11. The actual power drawn is calculated in the financial model.

Table 17.11 Summary of Connected Power for Heap-leach and Mill Operations

Area Description Connected Load, kW
HEAP-LEACH OPERATIONS
100 Primary Crushing 1,607
200 Fine Crushing 5,798
250 Agglomeration 436
300 Conveying and Stacking 3,158
350 Heap Leaching 6,867
400 Merrill-Crowe 1,795
500 Refinery 368
650 Water Systems 795
800 Reagent Facility 25
900 Compressed Air 57
Total Heap Leach 20,907
NON-OXIDE ORE MILLING OPERATIONS
120 Primary Crushing (Transfer Conveyors) 1,141
220 Crushed Ore Reclaim 204
320 Grinding 7,094
420 Flotation and Regrind 2,309
440 Concentrate Leaching 259


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Area Description Connected Load, kW
460 Concentrate CCD 419
620 Flotation Tailing Thickening and Handling 498
670 Water Systems 297
820 Reagents 252
Total Mill 12,473

17.11Control Systems

A crusher control room in the primary crusher area will be the operating and control center for the crushing plants, particularly for the primary crusher and Railveyor system. A central control room (“CCR”) will be located inside the Merrill-Crowe Plant building. From the CCR control consoles, crushing, screening, material handling systems, reagents, pumping systems, aeration system, Merrill-Crowe plant, and utility systems will be monitored or controlled.

A computer room adjacent to the CCR will contain engineering workstations (“EWs”), a supervisory computer, historical trend system, management information systems (“MIS”) server, programming terminal, network and communications equipment, and documentation printers. This will be primarily used for distributed control system (“DCS”) development and support activities by plant and control systems engineers.

Although the facilities are normally controlled from the CCRs, local video display terminals are to be selectively provided on the plant floor for occasional monitoring and control of certain process areas. Any local control panels that are supplied by equipment vendors will be interfaced with the DCS for remote monitoring or control.

17.12Assay and Metallurgical Laboratories

A laboratory building has been provided for in the capital cost estimate. Provision has been made for facilities that include sample receiving, sample drying, sample preparation, metallurgical laboratory, wet laboratory, and fire assay for mine and process plant samples.

17.13Alternative Processing Options

Integra has considered several processing tradeoffs in terms of grind size, processing rates, etc., with and without the use of a high-pressure grinding roll (“HPGR”) circuit. Also, the addition of flash flotation and cleaner flotation in the non-oxide plant are being considered for future installation. Finally, oxidation of flotation concentrates using the Albion process is being studied for inclusion in future technical reports.

17.13.1High-Grade Heap Leaching Ore Processing

A tradeoff study was conducted to determine if high-grade heap-leach ore would justify a finer crush using three stages of crushing with agglomeration. The third-stage crusher would either be a cone crusher or an HPGR. In addition, the study also examined the viability of sending a portion of the high-grade heap-leach ore to the mill.

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The parameters for the study are shown in Table 17.12 in the order of intensity of treatment. Table 17.12 includes the capital cost, operating cost, and metal recoveries for each option. For the tradeoff study, the capital and operating costs were based primarily on historical benchmarks and M3 Engineering’s in-house database information. The costs in the tradeoffs are exclusively intended for comparison within the tradeoff.

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Table 17.12 Processing Options for High-Grade Heap-Leach Ores

Option HL Mill HL Crushing Capex Process Ore Florida Mountain DeLamar
mtpd mtpd Circuit $M Au Rec, % Ag Rec, % Opex, $/t Au Rec, % Ag Rec, % Opex, $/t
1 30,000 0 2-Stage Cone $129.0 Heap Leach Oxide 90% 40% $2.55 85% 24% $2.65
50 mm Trans 80% 30% $2.85 66% 22% $3.60
2 30,000 0 3-Stage Cone $166.6 Heap Leach Oxide 90% 65% $2.90 85% 30% $3.00
Agglomerated

12.7 mm
Trans 85% 55% $3.55 70% 30% $4.00
3 30,000 0 3-Stage HPGR $172.4 Heap Leach Oxide 90% 69% $3.20 85% 36% $3.35
Agglomerated

6.4 mm
Trans 85% 60% $3.95 72% 34% $4.40
4 25,000 5,000 2-Stage Cone $266.6 Heap Leach Oxide 90% 40% $2.65 85% 24% $2.75
50 mm Trans 80% 30% $3.00 66% 22% $3.70
Mill Oxide 93% 80% $14.40 88% 67% $14.70
Trans 88% 89% $25.40 75% 58% $16.05
5 25,000 5,000 3-Stage Cone $304.2 Heap Leach Oxide 90% 65% $3.00 85% 30% $3.10
Agglomerated Trans 85% 55% $3.65 70% 30% $4.10
12.7 mm Mill Oxide 93% 80% $14.40 88% 67% $14.70
Trans 88% 80% $15.40 75% 58% $16.05
6 25,000 5,000 3-Stage HPGR $307.6 Heap Leach Oxide 90% 69% $3.35 85% 36% $3.45
Agglomerated Trans 85% 60% $4.05 72% 34% $4.55
6.4 mm Mill Oxide 93% 80% $14.40 88% 67% $14.70
Trans 88% 80% $15.40 75% 58% $16.05


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17.13.2Milling Options for DeLamar Non-oxide Ore

The base case scenario for milling non-oxide ore is to produce a flotation concentrate, fine grinding the concentrate and leaching the ground concentrate with cyanide. Options were studied to improve recoveries for DeLamar ores. These are shown in Table 17.13, which includes capital cost, operating cost, and metal recoveries for each option.

The first set of trade-off studies tested the base-case scenario at different mill throughputs, namely 5,000, 8,000 and 10,000 tonnes per day (Options 1 through 3). The second set of trade-off studies adds a leaching facility to leach flotation tailing (Options 4 and 5), at 5,000 and 8,000 tonnes per day. The third set of trade-off studies (Options 6 and 7) does not include a tailing leach facility but preoxidizes the concentrate using the Albion process at the finer grind of 10 microns.

Table 17.13 Milling Options

Option Tonnage Grind Tailing Albion Capex Domain DeLamar
mtpd P80,µ Leach $M Au Rec, % Ag Rec, % Opex, $/t
1 5,000 20 No No $135.0 Sullivan Gulch 45% 72% $12.52
Glen Silver 24% 64% $12.52
2 8,000 20 No No $176.6 Sullivan Gulch 45% 72% $11.30
Glen Silver 24% 64% $11.30
3 10,000 20 No No $203.2 Sullivan Gulch 45% 72% $11.02
Glen Silver 24% 64% $11.02
4 5,000 20 Yes No $186.2 Sullivan Gulch 45% 75% $17.96
Glen Silver 29% 71% $17.96
5 8,000 20 Yes No $250.7 Sullivan Gulch 45% 75% $15.62
Glen Silver 29% 71% $15.62
6 5,000 10 No Yes $170.6 Sullivan Gulch 84% 84% $20.06
Glen Silver 74% 74% $18.41
7 8,000 10 No Yes $226.4 Sullivan Gulch 84% 84% $18.66
Glen Silver 74% 74% $17.01

Leaching of the flotation tail for DeLamar ores did not improve gold recovery and only slightly improved silver recovery. In contrast, pre-oxidation using the Albion process almost doubled gold recovery and significantly improved silver recovery.

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17.13.3Non-oxide Ore Gravity Concentration

One option that was considered as a future installation is a gravity concentrator to process a bleed from the ball mill cyclone underflow. The gravity concentrate would be sent directly to concentrate regrind. The theory behind this option to is recover gravity-recoverable gold (“GRG”) and silver to possibly increase recovery by catching gold and silver that could be missed by flotation.

17.13.4Cleaner Flotation Stages

Cleaner flotation stages may be added in the future to improve concentrate grade and reduce the volume of concentrate that needs to be reground and leached. Tailing from the cleaner section will be returned to the rougher bank, which would increase overall flows because of the recycle streams. It is possible that concentrate grades, recoveries and mass pulls may be sufficient that the cleaner banks will not be necessary. The addition of cleaner flotation stages may be called for if the Albion Process would be included and the sulfide sulfur grade is not high enough (10% sulfide sulfide) to provide the required heat for the process.

17.13.5Process Personnel and Staffing

Staffing requirements for process personnel have been estimated by M3 based on experience with similar-sized operations in the region. Total process personnel requirements are estimated at 41 persons for the heap-leach and Merrill-Crowe operation. An additional process personnel of 49 persons would be required for the non-oxide mill circuit.

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18.0PROJECT INFRASTRUCTURE

The infrastructure for the DeLamar project has been developed to support mining and processing operations. This includes the access road to the facilities, power supply, Railveyor, communication, heap-leach pads, process plant, and ancillary buildings. Haul roads within the mining area as well as the mine waste storage facilities are described in Section 16.0.

18.1General Arrangement and Site Access

Figure 18.1 shows the DeLamar and Florida Mountain general arrangement drawing. This includes pit designs, waste-rock storage facilities, heap-leach pads, mill facilities and tailing storage facilities.

The main access to the project is via gravel roads from Jordan Valley, Oregon, as used for previous mining at DeLamar. The existing DeLamar site access road is located on the east side of Henrietta Ridge extending from the DeLamar Road across Jordan Creek to the western side of the existing reclaimed Kinross tailing impoundment (Figure 18.1). This existing site access road is expected to become unusable due to its proximity to the proposed Milestone pit haul road and DeLamar West WRSF. Therefore, this PFS proposes relocating the site access road to the west side of Henrietta Ridge.

The alignment of the proposed site access road extends from the DeLamar Road approximately 800 meters (2,625 feet) east of the junction of Trout Creek Road and DeLamar Road, 3.9 kilometers (2.5 miles) to the end of the existing site access road. The proposed site access road crosses both state and federal lands. The first 1.5 kilometers (0.9 miles) of the proposed site access road gradually descends from the DeLamar Road at a maximum grade of 8.0% to the Jordan Creek valley where Jordan Creek is crossed via a proposed 18.29-meter (60.0-foot) bridge. From the Jordan Creek bridge, the proposed site access road steadily ascends the remaining 2.4 kilometers (1.5 miles) around the southern end of Henrietta Ridge to the mine site at a maximum grade of 8.5%.

The proposed site access road is similar in geometry and design to the existing site access road. The proposed typical road section consists of two 3.66-meter (12.0-foot) lanes with up to two 0.91-meter (3.0 feet) shoulders on either side and will be gravel surfaced. The site access road will have a 0.61-meter (2.0-foot) high safety berm on fill-slope sides and a roadside ditch and 1:1 or 2:1 (horizontal:vertical) cut slope for cut-slope sides. The roadside ditches will be periodically dewatered via corrugated metal culverts. The proposed site access road was designed to meet all geometry requirements of AASHTO Guidelines for Geometric Design of Very-Low Volume Local Roads, Second Edition and/or A Policy on Geometric Design of Highways and Streets, 7th Edition.

The estimated cost of the proposed site access road and all associated structures, including the Jordan Creek Bridge, is estimated at $8.96 million. This estimate includes a gross bank cut of 253,015 cubic meters (330,944 cubic yards), and a gross fill of 76,745 cubic meters (100,382 cubic yards) for a net excess bank cut of 176,270 cubic meters (230,561 cubic yards) to be utilized elsewhere. The cost estimate includes an estimated 126,508 cubic meters (165,472 cubic yards) of drilled and blasted rock which contributes to a significant portion of the overall cost (approximately 35%). This PFS assumes that the construction of the site access road will begin in month -20 and will be completed within one 6-month construction season.

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Haul road access between the DeLamar mine and Florida Mountain will need to be improved for use with the proposed mining equipment. This access will be utilized for delivery of all consumables, as well as any required construction materials and equipment. This will also be the primary access for all personnel working at Florida Mountain.

Due to the high altitude and amount of snowfall during winter months, the access roads will be maintained using the mine graders along with a sanding dump truck fitted with a blade for snow removal.

Security fencing will be constructed around the facilities. Buildings will include a site warehouse, administrative offices, an assay lab, and a mine shop.

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Figure 18.1 PFS General Arrangement Drawing

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18.2Heap-Leach Pad Construction

The heap-leach pads (“HLP” or “HLPs”) area is shown in Figure 18.1 and will be located immediately north of the crushing facility in portions of Sections 3, 4, 9 and 10, Township 5 South, Range 4 West. The site slopes northerly toward Jordan Creek at an average gradient of 12.5 percent. The HLPs will be constructed in two phases. The Phase 1 portion (Figure 18.2) will be constructed on a feature locally identified as Jacobs Ridge and into an adjacent valley to the west (herein referred to as the “unnamed gulch” or the “valley”). The site is generally underlain with a basalt which is overlain with a thin veneer of colluvium derived from weathering of the basalt and interbeds of tuff. Upper portions of the HLPs are underlain with porphyritic latite lava flows. The northern extent of the Jacobs Ridge pad area is underlain by a Miocene age rhyolite dike or plug. Geotechnical drilling in the Jacobs Ridge portion of the site in 1988 identified discontinuous layers of weathered tuff that had low shear strength. An initial auger drilling program on the western side of the site did not encounter the tuffaceous material encountered on Jacobs Ridge.

Initial construction of Phase 1 will include growth media removal, construction of a haul road from Jacobs Ridge to the lower end of the valley, constructing a French drain in the valley floor, and excavation of approximately 2.5 million cubic meters (3.3 million cubic yards) of soil and rock from the ridge and placing it as compacted fill in the gulch. This extensive regrading flattens the HLP foundation for stability considerations and will provide platforms for solution collection ponds on Jacobs Ridge and at the north end of the gulch. During the regrading, any weak material encountered will be removed from the ridge and placed in the bottom of the valley where it will be confined and will not negatively impact stability of the heap. Following regrading, the Phase 1 pad area and solution ponds will be lined in accordance with the IDEQ Rules 58.01.13 – Rules for Ore Processing by Cyanidation. In accordance with the regulation, the lining system will consist of 0.6 meters (24-inches) of compacted clay overlain with an 80-mil thick high-density polyethylene (“HDPE”) liner – or approved equivalent. The general layout of the HLP and solution ponds is shown in Figure 18.2.

The solution ponds are designed to contain pregnant solutions that drain by gravity from the crushed ore stacked on the HLP plus direct precipitation onto the lined portion of the pad area. The ponds will store the runoff from a storm event that has a 1% probability of exceedance (100-year, 24-hour precipitation event) plus 0.6 meters (2.0-feet) of freeboard. There will be one operating solution pond immediately north of the Jacobs Ridge heap with a capacity of 71,540 cubic meters (18.9 million gallons) and two ponds at the north end of the valley heap (Figure 18.2). Due to the smaller lined area at Jacobs Ridge, the one pond will function as both the operating pond and as the event pond with 15,140 cubic meters (4.0 million gallons) of pregnant solution storage as the normal operating volume and sufficient capacity remaining in the pond to contain the 100-year event plus the required freeboard. The operating pond at Jacobs Ridge will discharge only by pumping to the process plant.

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Figure 18.2 Heap-Leach Pad General Arrangement

The operating pond for the valley pad will be a double-lined pond with a capacity of 45,000 cubic meters (11.9 million gallons) and a single-lined event pond with a capacity of 71,160 cubic meters (18.8 million gallons). The operating pond will contain process solutions and will overflow into the event pond following large precipitation events. Additionally, two lined ponds are provided upstream of the valley portion of the HLP for short-term storage of storm runoff and excess process solution. Surface runoff will be diverted around the HLP with temporary ditches between Phase 1 and Phase 2 construction and permanent diversion ditches above Phase 2 construction.

Following grading and installation of the lining system, the HDPE-lined surfaces will be covered with a minimum 1.3-meters (4.3 feet) thick layer of crushed ore to protect the liner from equipment loading in excess of 41.4 kPa (6 pounds per square inch). A network of perforated HDPE solution collection pipes will be installed simultaneously with the placement of the overliner ore. This pipe system conveys leach solutions out of the heap during operations. Ore will be loaded onto the pad in nominal 10-meter (32.8-foot) lifts, primarily with moveable conveyors. However, trucks may be used to assist with ore placement when relocating conveyors. The lifts will be started at the north end of the pad and will extend southerly with a level top surface. As each new lift is started, it will be set back approximately 15 meters (49.2 feet) from the crest of the previous lift so that the final overall slope of ore on the pad can be graded to a 3H to 1V (18.3 degrees) for reclamation. A cross section of the loading geometry is shown in Figure 18.3.

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Figure 18.3 Heap-leach Ore Stacking Plan Geometry

X

Phase 2 portion of the HLP will consist of a westerly extension of the pad and tying in the area between the west side of the Jacobs Ridge pad and the east side of the Phase 1 valley pad. Construction of Phase 2 will begin two years ahead of when the extended pad is needed, assumed in year 3 of operation. Phase 2 construction will be performed in the same sequence of activities and will add approximately 30% to the pad footprint. The total volume of ore to be placed on the HLP is between 95 million tonnes and 100 million tonnes which may include up to 2 million tonnes placed at the southern end of the Jacobs Ridge portion of the Phase 1 pad to minimize recovery time from the final ore placed on the pad.

18.3Slaughterhouse Gulch Tailing Storage Facility

The primary flotation tailing disposal facility (“TSF”) for the DeLamar project will be located in Sections 30 and 31, Township 4 South, Range 4 West, and Sections 25 and 36, Township 4 South, Range 5 West, in Slaughterhouse Gulch, approximately 6.0 kilometers (3.7 miles) west of the new mill site (Figure 18.1). Slaughterhouse Gulch is a natural drainage that descends to the south primarily on State and BLM lands. The TSF will be a zoned earth and rockfill embankment that will be located where the valley narrows approximately 1 kilometer (0.6 miles) north of its confluence with Jordan Creek. The Slaughterhouse Gulch TSF will impound flotation tailing that have not been processed by cyanidation and therefore will not be lined in accordance with IDEQ 58.01.013. The earth dam will be designed in accordance with Idaho dam safety regulation IDAPA 37 – DEPARTMENT OF WATER RESOURCES Water Allocations Bureau 37.03.05 -Mine Tailings Impoundment Structures.

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Slaughterhouse Gulch is covered with a veneer of alluvium overlying a dense tuffaceous clay layer of variable thickness from 3.05 meters to 12.19 meters (10 feet to 40 feet), which overlies a weathered basalt bedrock. The depth to groundwater is variable from 3.05 meters to 6.1 meters (10 feet to 20 feet) – generally within the clay layer or between the clay and bedrock. Falling head tests in auger drill holes indicate a low transmissivity value in the clay and basalt.

The TSF will be constructed in stages in a downstream sequence. The initial “starter” embankment will be constructed with one or more zones using borrow material from within the impoundment area as fill. Earth fill will be compacted in layers usually not exceeding 35 centimeters (13.8 inches). Rock fill will be placed in lifts up to 152.4 centimeters (60.0 inches) thick and compacted with dozers and mine haulage equipment. Portions of rock fill in the downstream “shell” zone will primarily be waste-rock from a nearby open pit. The starter embankment will be 60 meters (197 feet) in height and will contain sufficient storage for approximately three years of mill production plus freeboard for storage of stormwater runoff. Subsequent stages of embankment construction will be performed to raise the elevation of the embankment to provide additional tailing storage. The centerline of the embankment will move downstream as additional fill is placed, so that new construction will not be placed on tailing material. The fill for subsequent stages will come from a combination of borrow material from upstream areas of the impoundment and from mine waste rock. The ultimate crest height of the embankment will be 86 meters (282 feet), which will allow storage of up to 23 million tonnes without bottom drainage, or 26 million tonnes of flotation tailing with bottom drainage, based on laboratory column settling tests. The final configuration of the TSF will encompass approximately 97 hectares (240 acres).

Thickened tailing will be deposited from spigots upstream of the embankment as a slurry. The slurry will form a beach in front of the embankment as the coarser particles settle and a process water pool will form away from the embankment where the finest particles will settle from the slurry. The spigots will be managed to control dust, the beach surface, and the water pool location. Cyanide-free decant water will be collected at the eastern edge of the impoundment and will be pumped back to the process water tank for the use in the front end of the mill. A minor amount of seepage will flow through the embankment and will be collected in a lined pond, and then will be pumped back into the impoundment.

Runoff from normal precipitation events occurring above the ultimate impoundment footprint will be diverted around the TSF with a diversion ditch that will carry these intermittent flows southwesterly to a discharge point on the west side of the impoundment. The diversion ditch will carry up to a 10-year, 24-hour runoff event. Storms larger than a 10-year event, up to a 100-year, 24-hour event, will be stored temporarily on the TSF impoundment and used as makeup water for the mill. Larger storm events will be discharged through a spillway provided for each stage of the embankment to protect the embankment from overtopping. The final stage of the embankment will be protected from overtopping with a spillway that will safely carry the largest event mandated by state and federal agencies having jurisdiction over dam safety issues in view of potential changes in climatic conditions.

The embankment will be constructed according to accepted standard design principles developed by the U.S. Corps of Engineers and Bureau of Reclamation for water retaining dams, except the design will allow the entrained fluids to seep through the dam in a controlled fashion so that the tailing solids can consolidate. A cross section of the conceptual design is provided in Figure 18.4.

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Figure 18.4 Schematic Cross Section of Tailing Impoundment Embankment

Note: numbers in circles refer to “zones” defined below.

The embankment zones shown with numbered circles in Figure 18.4 are as follows:

  • Zone 1 is the core of the dam that is constructed with compacted clayey soil to limit seepage. The Zone 1 core will be extended down into a shear key trench that will be provided to intercept existing natural preferential flow paths that may exist under the embankment footprint;

  • Zone 2 is used as upstream and downstream fill to contain the core of the embankment. The Zone 2 fills are typically higher shear strength soils containing sands and gravels with some clay;

  • Zone 3 is a chimney drain to allow seepage through the core to be transmitted vertically to an exit drain under the downstream shell of the embankment for stability enhancement;

  • Zone 4 is an upstream blanket drain that will allow the tailing to consolidate. This zone will be compacted sand and gravel and will drain vertically and be evacuated periodically with a well pump to assist the settled tailing solids layers to drain as additional layers are placed during operation; and

  • Zone 5 is the downstream shell of the embankment. The shell provides mass to the embankment that resists movement as the impoundment is filled. It will be constructed with waste rock from an adjacent mine pit and from quarried rock. The Zone 5 rock will be compacted in lifts up to 1.5 meters thick with dozers and mine haulage equipment.

The embankment design anticipates 2H to 1V upstream slope and 2.5H to 1V downstream slopes. Stability analyses based on strength parameters determined by triaxial testing on hollow-stem auger samples collected during a 2021 site investigation indicate a static factor of safety of 1.66 and a pseudo-static (earthquake) factor of safety of 1.24 of the downstream embankment slope (Figure 18.5). The stability of the upstream slope will be higher than the downstream slope due to the gradient of the foundation and will be immediately enhanced when tailing is placed in the impoundment. The final design will be based on closer spaced drilling and test pit sampling to provide additional confirmation of geotechnical strength and consolidation characteristics of foundation soils. There is adequate room to flatten the downstream slope if necessary to provide adequate factors of safety following additional investigation.

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Figure 18.5 Embankment Components Cross Section

Prior to embankment construction and impoundment filling, growth media will be removed to an average depth of 0.3 meters and placed in one or more stockpiles outside of the maximum footprint of the facility. The locations of the stockpiles will be determined following final design. This material will be placed back onto the surface of the tailing during reclamation and closure.

18.4Concentrate Leach Tailing Storage Facility

The concentrate leach tailing storage facility (“CLTSF”) will be a smaller, 26 hectare (64.2 acre) impoundment for containment of flotation concentrates from the milling process after they have been leached with cyanide to remove precious metals. To aid in settling, this fine material (P80 of 20 microns) will be blended with a small stream of coarser flotation tailing in roughly a 1:1 blend. The location of this CLTSF is immediately south of the HLP at the head of the unnamed drainage (Figure 18.1). The construction of the CLTSF in this location will involve placing fill from the Jacobs Ridge pad area to provide initial stormwater storage and then installing a liner system in year 2 that will meet the lining requirements of the IDEQ Rules 58.01.13 – Rules for Ore Processing by Cyanidation. In accordance with the regulation, the lining system will consist of 61 centimeters (24 inches) of compacted clay overlain with an 80-mil thick HDPE liner – or approved equivalent. The downstream side of the TSF will be constrained by crushed ore placed in the south end of the HLPs. A geotextile will be placed on the ore to allow drainage from the CLTSF into the ore to enhance consolidation of the tailing during operation and following closure. Excess fluids will be decanted from the surface of the impoundment and pumped back to a tank for re-introduction into the process water stream. Since this impoundment will be constructed in accordance with the IDEQ Cyanide Rules, in may also be used for temporary storage of excess fluids containing cyanide due to precipitation events on the HLP.

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The southern side of the CLTSF will be the natural drainage divide between the unnamed gulch to the north and Sullivan Gulch to the south (Figure 18.1). The natural ridge will be raised to a sufficient elevation to provide the required impoundment capacity. The resulting embankment will be designed to State Engineer requirements for stability and storm runoff containment. The impoundment side of the embankment will be lined in accordance with IDEQ 58.01.13. Surface runoff from small upgradient watersheds east and west of the impoundment will be diverted around the impoundment with small drainage ditches.

The maximum tailing depth is anticipated to be 51 meters (167 feet). The projected capacity of concentrate tailing is approximately 5.2 million tonnes at this depth. The conversion from a stormwater management facility will involve 1) completion of the clearing and foundation, 2) preparation of the impoundment and southern embankment area, 3) extending the valley French drain to intercept springs and seeps under the CLTSF, 4) placing compacted fill in the bottom of the valleys to provide a suitable platform for installing the lining system, 5) constructing the southern embankment, and 6) installing the lining system. Suitable leak detection and collection systems will be installed during construction. Crushed ore will be placed to raise the HLP side of the embankment as necessary to provide containment for the concentrate leach tailing and precipitation inflows. Further geotechnical investigations are scheduled in 2022.

18.5Renewable Energy Systems

The DeLamar project power supply has been planned using renewable energy systems along with a liquified natural gas (“LNG”) plant in order to reduce emissions for the project. Work on these power sources was provided by John F. Gardner of Warm Springs Consulting in Boise, Idaho.

18.5.1Power Generation and Distribution

The electrical power demand at the DeLamar project facilities is currently estimated at 13.5 MW for initial heap-leach process operations, with an additional load of 9.8 MW for the mill circuit. The demand will vary according to the quantity of each ore type to be processed. The average load for the mine is forecast to be 11.6 MW (Table 18.1) with a peak demand of 23.4 MW. Lifetime electricity consumption is estimated to be 1.8 million MWh.

Existing electrical infrastructure on the project site consists of a 69 kV transmission line operated by Idaho Power Company. Significant upgrades to existing electrical infrastructure would be required to meet the anticipated load increase associated with the project, including construction of new 138 kV transmission lines, substations and tap station upgrades. To reduce capital expenditures of energy infrastructure, ensure power supply resilience and reduce emissions, Integra plans to power the project through an on-site microgrid with a solar electrical generation system and an LNG plant.

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This microgrid is planned to consist of a 12 MW solar array that will be installed on the historical tailing impoundment in conjunction with 4.5 MWh of batteries (Table 18.1). An LNG power plant is planned to be constructed on site and leased from a third-party provider through a long-term use-based equipment lease.

Carbon emissions from this energy mix will be an estimated 13% lower than using the current local utility grid mix. While solar will meet an estimated average of 19% of annual energy consumption, Integra will continue to look at how to affordably increase the percent of energy provided by renewables while ensuring energy reliability. Table 18.1 shows the power consumption and power production estimates.

The photovoltaic (“PV”) solar array will be located on the legacy tailing impoundment area (Figure 18.1) which has around 121 hectares (299 acres) of available surface area. Battery backup will provide excess capacity during peak-demand periods and generator ramp-up, enabling operational flexibility. Natural gas will be transported to site on high-pressure mobile natural gas storage trailers.

Table 18.1 Power Consumption and Production Details

Parameter Units Quantity
Project Average Hourly Consumption MWh 11.66
Photovoltaic Array Sizing (PV) MW 12
Battery sizing MWh 4.5
Land use solar hectares 27
Average Annual power consumption kWh 102,118,941
Average Annual Solar Production kWh/yr 19,817,603
Average annual renewable fraction % 19%
Average Annual Generator Production kWh 82,301,338
Average % of energy from Generator % 81%

18.5.2Power Pricing

In 2020, the average levelized cost per MWh for contract solar projects nationwide was $24/MWh ($0.024 per kWh) without batteries, and around $40/MWh ($0.04 per kWh) for 90% battery to PV capacity, according to market studies. Based on research conducted by the study team via interviews with utility-scale microgrid developers in the region, a microgrid cost for this site could range from $0.05 to $0.07 per kWh, depending on the renewable fraction. The cost per kWh used for the cash flow model is $0.065 per kWh.

A trade off study was completed to assess capital and operational costs for the microgrid scenario relative to costs associated with upgrading the Idaho Power Company transmission infrastructure to site. The analysis considered both owner-operated on-site electrical generation and a third-party long-term purchasing agreement. The microgrid levelized cost of energy (“LCOE”) is estimated at $0.055/kWh, 63% lower than the local electric utility LCOE of $0.149/kWh, which included both the power rate and the capital expense associated with the power line upgrade. With a carbon price of $50/tonne of CO2 added in, the LCOE for the microgrid is $0.065/kWh and $0.168/kWh for grid supplied power.

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18.6Railveyor Haulage System

The project will utilize a Railveyor light rail haulage system to transport ore from the open pits to the crusher facility. The Railveyor system is an autonomous materials haulage system consisting of transport trains, light-rails, electrical drive stations, and materials loading and discharge stations. The system functions similar to a conveyor, but is designed to be modular and relocatable, allowing improved operational flexibility and lower cost. By leveraging the Railveyor system, the DeLamar project has a unique opportunity to realize cost savings compared to typical truck haulage, while reducing its overall fuel consumption and carbon footprint and automating many essential functions that typically would require on-site personnel.

Integra conducted an internal cash flow tradeoff study that evaluated 136-tonne diesel haul trucks vs Railveyor for haulage of ore from designed pit loadout areas to the mill. Incorporating Railveyor will replace approximately five haul trucks, reducing fuel consumption from haulage by approximately 20% with an overall reduction of diesel consumption by 13%. Fuel and labor savings provide significant reductions in operating expenditures with increased capital costs. However, the entire project net present value (“NPV”) was higher with Railveyor versus trucks only. Additional environmental benefits will include reduced noise levels, reduced dust levels, and reduced water consumption for dust mitigation.

There will be two Railveyor systems in operation. The Florida Mountain system will be in operation during years 1 through 6. The DeLamar system will come online in year 4 for development of the DeLamar main pit. In essence, the system consists of a multi-train loop with approximately 19 kilometers (11.9 miles) of rail (Figure 18.1). Railveyor features a metal rail fixed atop gravel and which can be repositioned as the layout of the mine develops. Key parameters for the Railveyor design are summarized in Table 18.2. Ten trains in each system would be capable of transporting 2,100 to 2,700 tph and would consist of approximately 150 individual cars with a total length of 340 to 380 meters (1,116 to 1,247 feet). Trains will travel from loading points near pits along the haul roads (Figure 18.1) on a dedicated track to a single discharge loop. Trains will proceed around an inverted discharge loop which empties cars into the crusher dump pocket. Haul trucks will be utilized to transport material from shovels to the start of the Railveyor system.

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Table 18.2 Key parameters for Railveyor System

Parameter Florida Mountain DeLamar
System Design Rate 2,114 tph 2,718 tph
Load Rate 2,600 tph 2,800 tph
Train loading speed 1.18 m/s 1.10 m/s
Train tramming speed 3.0 m/s 3.0 m/s
Train return speed 4.0 m/s 4.0 m/s
Distance – load point to mill discharge 5,305 m 4,001 m
Total cycle time per train 64 minutes 51 minutes
Trains required in system 10 10
Train length 383 m 339 m
Cars per train 157 cars 139 cars
Material weight hauled per train 232 tonnes 237 tonnes
System track length (rails) 11,280 m 8,010 m
Total Drive Stations 81 55

The Railveyor trains will be powered by electric drive stations placed along the track throughout the length of the system. Drive stations consist of 100 hp AC motors geared to commercial truck tires which squeeze the rail cars to propel and stop the train. Because the system will use electric drives to propel rail cars, it will be powered entirely by the onsite electrical system. Railveyor also allows for electricity generation through regenerative braking which results in a net surplus of power generated in the downhill-haul Florida Mountain loop.

18.7Project Buildings

The proposed heap-leach facility will be located between the DeLamar and Florida Mountain pits (Figure 18.1). The primary crusher and process facilities will be located just south of the HLPs. Ore will be conveyed from the primary crusher to oxide or non-oxide coarse ore stockpiles accordingly.

Oxide and mixed ore will be conveyed from the oxide coarse ore stockpile to the secondary and tertiary crushing and screening. Crushed oxide and mixed ore may report to the agglomeration circuit or bypass agglomeration if not required. From the agglomeration circuit, the ore will be conveyed to the leach pad via an overland conveyor to a series of grasshopper conveyors that will distribute the ore onto the pad in the prescribed courses. The grasshopper conveyors are not demonstrated on the drawings at this time for clarity as they will be moved throughout the loading of the heap-leach pad. The PLS will flow by gravity to the PLS ponds directly north of the heap-leach pad. An event pond will be located adjacent to the Valley PLS pond to allow for passive overflow if an excessive runoff event occurs. Road access is provided along the east edge of the heap-leach facility to allow access to the ponds. This access route will also serve as a pipe route for PLS and reclaim water piping to be pumped back up to the Merrill-Crowe and process facility pad.

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Non-oxide ore will be conveyed from the non-oxide coarse ore stockpile to the mill facility on the process facilities pad. Ore will be subjected to size reduction through the SAG mill and the ball mill. Primary ground ore will pass through one stage of rougher flotation. Concentrate from the flotation cells will be reground, thickened, and sent through a series of agitated leach tanks. PLS from the tank leach circuit will be recovered via CCD cyclone systems, with barren solution from the Merrill-Crowe circuit utilized as the wash solution. The PLS will be clarified in a clarifying thickener and sent to the Merrill-Crowe facility for metal recovery. Post CCD slurry will be thickened, blended to a 1:1 solid mass ratio with flotation tailing slurry, and deposited in the dedicated TSF near the Valley fill leach pad as summarized in Section 18.4. Slurry from the rougher flotation tailing (non-cyanide bearing) will be pumped from the flotation tailing thickener to the west along haul and mine roads to the Slaughterhouse Gulch TSF, approximately 13.3 kilometers (8.4 miles) west of the process facilities pad.

Other buildings located on or near the process facilities pad include the administration/change building, a substation, assay lab, Merrill-Crowe plant, and water treatment plant.

A truck shop is also planned south of the primary crusher and ROM pad (Figure 18.1). Light vehicle and diesel fuel islands will be constructed just east of the truck shop. A truck wash and tire change pad are also included southeast of the truck shop. Safety and training areas will be provided within the truck shop building. In addition, mine services offices are integral to the truck shop and a laydown yard is proposed directly southwest of the facility. The DeLamar, Milestone, Sullivan, and Florida Mountain pits will be connected to their respective waste dumps and the primary crusher by haul roads.

All facilities will include all necessary eyewash/safety shower water and fire protection systems where needed.

18.7.1Crushing Facilities

The process infrastructure includes three structures for the crushing circuit: primary crushing structure, secondary and tertiary crushing structure, and the tertiary crusher discharge transfer tower structure.

18.7.2Mill and Flotation Buildings

The mill and flotation buildings will be uninsulated, engineered steel buildings and have been designed to protect the mill and flotation components from harsh weather elements. The mill and flotation buildings will have walls on all four sides and include several roll-up doors to facilitate access by maintenance and support equipment. Each building will also include a separate bridge crane to be used for maintenance and daily activities. The mill building will be approximately 62.5 meters (205.0 feet) x 29.0 meters (95.1 feet) and have an eave height of 23 meters (75.5 feet). The flotation building will be approximately 76.0 meters (249.3 feet) x 27.5 meters (90.2 feet) and have an average eave height of 23 meters (75.5 feet).

18.7.3Security Building at Access Gate

The site security building will be located on the east side of Henrietta Ridge, approximately four kilometers (2.5 miles) along the main access road from Trout Creek Road. The security building includes an entry access gate that will control all site ingress and egress.

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18.7.4Administration And Changing Building

The site administration and changing building will be located on the process facility pad seven kilometers (4.4 miles) east of the access gate and security building. The building will be comprised of mobile units each of 3.7 meters (12.1 feet) x 18 meters (59.1 feet) that will be assembled into a single unit divided for the variety of use. Ten of these units will be used for administration and two will be used for the changing facilities.

18.7.5Truck Shop Building

The truck shop will be located just south of the primary crushing and ROM pad. It will have dimensions of 90 meters (295.3 feet) x 30 meters (98.4 feet) with six bays, two of them will utilize embedded rail to receive tracked vehicles or loaders with tire chains. The mine warehouse facility will be included within the footprint of the truck shop on the ground floor at the opposite of the bay side. The mine services offices and training space is designed to be included above the warehouse space.

18.7.6Truck Wash Building

The truck wash will be located adjacent to and south of the truck shop. The truck wash is a covered and enclosed 26-meter (85.3-foot) x 24.5-meter (80.4-foot) facility that will have one bay along with water tanks and pumps. The truck bay will be 15.2 meters (49.9 feet) x 24.5 meters (80.4 feet) and will be for the large haul truck fleet. The area adjacent to the truck bay will be a basin for solids and oil removal, along with an area for pumps, a water tank, and support equipment.

18.7.7Merrill-Crowe Plant and Refinery Buildings

The Merrill-Crowe plant will be located directly to the south of the heap-leach pad at the northwest corner of the process facilities pad (Figure 18.1). PLS from the heap leach will be processed in the Merrill-Crowe plant where gold and silver will be precipitated via zinc cementation and recovered as a filtered cake in the refinery. The Merrill-Crowe facility includes the clarifier filter area and single deaeration tower. The precipitate filters, which are downstream of zinc addition, will be located in the refinery building, which is a covered and enclosed 32.5-meter (106.6-foot) x 17-meter (55.8-foot) secure building with an eave height of 5.7 meters (18.7 feet). The refinery building will contain the precipitate filters, mercury recovery (retort), and smelting furnace. The refinery will include a secure man-door access as well as access for vehicles via a roll-up door.

18.7.8Laboratory Building

The assay laboratory building will be comprised of a series of mobile buildings that will be assembled into a single unit to allow for a more conventional layout. The layout will include six 3.7-meter (12.1-foot) x 22-meter (72.2-foot) buildings (18.3-meter (60.0-foot) x 22-meter (72.2-foot) building footprint) and accommodate proper scrubbers, acid containment system, dust collection, and necessary sample processing equipment. Offices, restrooms, and change facilities for the laboratory will be incorporated into the layout.

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18.8Water Management Systems

18.8.1Fresh Water

Water consumption for the process facilities is summarized in Section 17.8. For the project, it is anticipated that there will be several freshwater wells on-site that will provide the requirements of the project. Fresh water will be stored in a fresh/fire water tank that will have reserve storage dedicated for fire protection. The balance of the fresh/fire water volume will be utilized to supply the demands of the process as well as mine dust suppression. Figure 18.6 provides an overview of the fresh water points of use within the process facilities.

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Figure 18.6 Water Systems Flow Sheet

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18.8.2Water Treatment Plant

Stormwater from the site will be managed as contact and non-contact stormwater. Non-contact stormwaters are the flows that do not come in contact with ore or mine processing facilities. Non-contact flows will be diverted and conveyed around the sites and directly discharged to existing stream channels. Contact stormwater will be utilized within the process to the greatest extent that allows the process to maintain a neutral balance. If there is excess contact water within the process, the excess will be routed to a water treatment plant. There is an existing water treatment plant at the project site. An allowance has been included for additional water treatment capacity consisting of a plant with solids separation and treatment, as required, to allow for discharge to existing stream channels or re-use in the process system.

18.9Mine Site Personnel

Mine site personnel requirements are shown in Table 18.3. This includes administrative, mining, and processing. In addition, there would be approximately 80 additional personnel working on-site during construction.

Table 18.3 Mine, Process and Administrative Personnel

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19.0MARKET STUDIES AND CONTRACTS

19.1Metal Pricing

No market studies have been undertaken for this PFS. Gold doré will be the commercial product from the DeLamar and Florida Mountain operation. Gold doré is readily sold on the global market to commercial smelters and refineries, and it is reasonable to assume that doré from the project will also be salable.

To determine appropriate metal prices to be used for economic analysis and cutoff grades, Mr. Dyer has considered spot prices for gold and silver in the months prior to the effective date of this report and has reviewed current metal prices used in recent NI 43-101 technical reports. In addition, three-year and two-year rolling average gold and silver prices were reviewed along with one-year forward pricing.

As of the end of January 2022, the three-year rolling average gold price based on Kitco metal pricing was $1,669 per ounce. The three-year rolling average silver price was $20.83 per ounce. This compares to the two-year rolling averages of $1,795 and $23.05 per ounce for gold and silver, respectively, at the end of January 2022.

A review of 10 different technical reports from mid- to late-2021 showed gold pricing ranging from $1,500 to $1,750 per ounce gold. Thus, there does not seem to be a lot of consistency on gold prices used. Of these reports, only four projects included silver with ranges of $20.00 to $22.00.

Mr. Dyer has reviewed these studies and the rolling average and future prices and has settled on using a mix of consensus one-year future prices along with three-year rolling average prices. This yields a gold price of $1,700 per ounce and a silver price of $21.50 which have been used in the economic analysis discussed in Section 1.0.

Other than royalty contracts as discussed in Section 4, the QP is not aware of any other contracts in place at this time material to the user.

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20.0ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR COMMUNITY IMPACT

As discussed in Section 6.0 of this report, the DeLamar area has been mined extensively over the past century and most recently in the 1990s into early 2000s. Historical mining activities have altered the topography, hydrology and ecology of the area. Reclamation efforts undertaken at the site by Kinross have been conducted pursuant to federal and state permit requirements and have included waste-rock disposal area reclamation, tailing reclamation efforts, facility removal and cleanup, and surface disturbance reclamation. Integra has primarily conducted site permit compliance, water management activities and exploration since acquisition of the property in 2017.

20.1Environmental Studies and Permitting

The DeLamar project is located on public lands administered by the BLM, State of Idaho lands, and private lands controlled by Integra in Sections 25, 35, and 36, Township 4 South, Range 5 West (T4S, R5W), Sections 30 and 31, T4S, R3W, Sections 1 through 16, T5S, R3W, Sections 28 through 36 T4S, R4W, Sections 1 through 3, 10, 11, 14, 15, and 22, T5S, R5W, and Sections 1 through 16, T5S, R4W, Boise Base and Meridian. There are a few private land parcels within the DeLamar project that are not controlled by Integra and no project activities are planned on those parcels. The access to the DeLamar project is via the Jordan Creek Road from Jordan Valley, Oregon to the project area. In general, the proposed mine operations will consist of four open pits, three WRSFs, and two TSFs, and the processing of the ore will use heap leaching and milling methods. Integra plans the construction, operation, reclamation, and closing of this mining operation. Major components include:

  • Four areas of open pits;

  • Four waste-rock storage areas;

  • Crushing and conveying system;

  • A heap-leach processing facility;

  • Two tailing storage facilities

  • One mill processing facility;

  • Reagent area;

  • Laydown areas;

  • A water delivery and distribution system;

  • Onsite solar power and LNG power generation system;

  • A power delivery and distribution system;

  • Excess water management system, including a surface discharge, a land application, or underground injection;

  • Storm water diversion ditches and storm water sediment basins;

  • Haul roads;

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  • Upgrade of the existing access road to the project; and

  • Truck shop, warehouse, Merrill-Crowe circuit and refinery, fuel storage, and laboratory.

Integra proposes to mine approximately 35,000 tonnes per day of heap-leach and 6,000 tonnes per day of mill-grade mineralized material (heap-leach only in Stage 1 and heap-leach and mill in Stage 2). Stage 2 construction would commence in one year at Integra’s discretion. The LOM strip ratio will be 2.21 tonnes of waste per one tonne of ore. The life of the operation will be 16 years.

The mineralized material and waste would be extracted from the open pits using conventional surface mining methods of drilling, blasting, loading, and hauling. Integra would use hydraulic shovels or front-end loaders to load the blasted mineralized material and waste into the haul trucks. The haul trucks would transport the waste rock to the rock disposal area near the open pits. The haul trucks would also transport the mineralized material to a load out location, where a Railveyor system would transport the material to the crushing system where the mineralized material would be crushed and delivered to the heap-leach pad, or the mill, for processing using a NaCN solution to leach the precious metals. After milling processing, the residual ore would be placed in a TSF. A Merrill-Crowe process would be used to precipitate the precious metals. The precipitate would then be refined in a furnace to produce doré bars for shipment off site. The project facilities would disturb over 809 hectares (2,000 acres).

The review and approval process for the Plan of Operations by the BLM constitutes a federal action under the National Environmental Policy Act (“NEPA”) and BLM regulations. Thus, for the BLM to process the Plan of Operations, the BLM is required to comply with the NEPA and prepare either an Environmental Assessment (“EA”), or an Environmental Impact Statement (“EIS”). Based on discussions with the BLM, Integra anticipates an EIS will be required to comply with NEPA.

The following sections provide information on historical and recent site characterization efforts, existing environmental conditions, status of project approval and permitting efforts, social and community considerations, proposed mitigation of stream and wetland disturbance, and reclamation and closure activities.

20.1.1Environmental Baseline Studies

Integra has contracted qualified third parties to perform environmental adequacy reviews of all available existing environmental baseline reports and data compiled from 1979 through present. Additionally, an EA was approved in 1987 for the DeLamar Silver Mine and an EIS was approved in 1995 for the Stone Cabin Mine by previous operators for the site.

In 2020, Integra conducted a technical adequacy audit of all existing environmental information, and began the collection of data that included the following:

  • Surface water hydrology and quality;

  • Ground water hydrology and quality;

  • Geochemistry;

  • Water rights; and

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  • Geotechnical/engineering.

Baseline studies for surface water were initiated in spring of 2020 and ground water studies were initiated in the spring of 2020. Geotechnical investigations for site features commenced in 2021 and geochemical fieldwork and kinetic testing commenced in 2020 and will continue into 2022/2023.

In 2021, Integra developed plans of study that included the following:

  • Wetland Resources Baseline Plan of Study;

  • Aquatic Resources Baseline Plan of Study;

  • Wildlife Resources Baseline Plan of Study;

  • Class III Cultural Resources Baseline Plan of Study;

  • Groundwater Baseline Plan of Study;

  • Surface Water Baseline Plan of Study;

  • Vegetation Baseline Plan of Study;

  • Soils Baseline Plan of Study;

  • PM10 Quality Assurance Monitoring Plan; and

  • Geochemical Characterization Plan of Study.

In 2021 Integra, working closely with the BLM and state agencies, completed the review and approval of the initial environmental baseline work plans. Baseline surveys initiated in accordance with the 2021 plans of study and baseline technical reports are underway. Additional plans of studies and collection of data will be undertaken in 2022, these include the following:

  • Transportation and public safety;

  • Recreation and visual;

  • Air quality/noise;

  • Hazardous materials;

  • Reclamation; and

  • Socioeconomics.

The data collection and technical reports are scheduled to be completed in the second half of 2022. The entire DeLamar mining district has been studied extensively, both historically and currently; therefore, ensuring scientific integrity of the methodologies and analysis used to collect the data and ultimately a meaningful analysis would be conducted allowing for a reasonable comparative assessment of the alternatives.

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20.1.2Federal Permitting

20.1.2.1Bureau of Land Management Plan of Operations

The Mine Plan of Operations (“MPO”) is submitted to the BLM for any surface disturbance in excess of five acres (2.02 hectares). The MPO describes the operational procedures for the construction, operation, and closure of the project. As required by the BLM, the MPO includes a waste-rock management plan, quality assurance plan, a storm water plan, a spill prevention plan, reclamation plan, a monitoring plan, and an interim management plan. In addition, a reclamation report with a Reclamation Cost Estimate (“RCE”) for the closure of the project is required. The content of the MPO is based on the mine plan design and the data gathered as part of the environmental baseline studies. The MPO includes all mine and processing design information and mining methods. The BLM determines the completeness of the MPO and, when the completeness letter is submitted to the proponent, the NEPA process begins. The RCE is reviewed by BLM and the bond is determined prior to the BLM issuing a decision on the MPO.

The MPO will be submitted for the project when operational and baseline surveys are complete and operations and design for the project are at a level where a MPO can be developed to the necessary level of detail. Submittal of the MPO is likely to occur in the first half of 2023.

20.1.2.2U.S. Army Corps of Engineers Authorization, Section 404 of Clean Water Act

A permit under Section 404 of the Clean Water Act (“CWA”) is required for the discharge of dredged or fill material into waters of the U.S. (“WOTUS”). Dredged or fill material includes tailing, heap-leach facilities, and waste rock. Other activities, in addition to the tailing and waste-rock storage that may require a Section 404 Permit are:

  • Road construction;

  • Bridges;

  • Construction of dams for water storage;

  • Stream diversions; and

  • Certain reclamation activities.

WOTUS include wetlands. A 2010 U.S. Supreme Court decision found mine tailing to be “fill,” and can, therefore, be placed in WOTUS with an approval under Section 404 of the CWA.

20.1.2.3Environmental Impact Statement

Approval of any MPO and Reclamation Plan by the federal agencies for the project as well as accordance with Section 404 requires an environmental analysis under the NEPA. NEPA requires federal agencies study and consider the likely environmental impacts of the proposed action before taking whatever federal action is necessary for the project to proceed.

The purpose and need for the project would be to conduct open pit mining and ore processing, which would disturb over 809 hectares (2,000 acres) of unpatented and patented mining claims and state lands within the project area and complete reclamation and closure activities, as well as long-term water treatment, to produce silver and gold from mineralized material of the estimated mineral resources. As a result, Integra anticipates that an EIS will be required to meet agency NEPA requirements.

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The BLM will be the lead federal agency for the preparation of the EIS, and other agencies will be cooperating agencies. The EIS and associated Record of Decision (“ROD”) effectively drives the entire permitting process timeline.

20.1.3Idaho Permitting

20.1.3.1Idaho Pollutant Discharge Elimination System Permit (“IPDES”)

An IPDES Permit would be required for point source discharges from the mining operation to WOTUS. Likely point discharges would include treated mine drainage, treated net precipitation from the TSFs, and any other discernible or discrete point source associated with mining and processing at the site. In addition, the project would be subject to performance standards for new sources for its respective industrial source category. This means the project would have to demonstrate that it is applying the best available control technology to meet applicable water quality standards. The permit application must be submitted at least 180 days prior to the approved discharge.

Storm water discharges associated with this industrial activity require a related permit. Storm water is defined as “storm water runoff, snowmelt runoff, and surface runoff and drainage.” Where flows are from conveyances that are not contaminated by contact with overburden or other mine waste, a permit is not required. Hence, the water management scheme developed for the project endeavors to collect and convey clean water around the mining operation and discharge downstream. Active storm water would be managed via a storm-water pollution prevention plan, which must also be submitted at least 180 days before commencing the discharge.

20.1.3.2Other Major State Authorizations, Licenses, and Permits

The key authorizations, licenses, and permits required by the State of Idaho are summarized in this section. The federal and state application processes would be integrated and processed concurrent with the EIS as follows:

  • Air Quality Application for Permit to Construct and Operate – Assesses the allowable impacts to air quality and prescribes measures and controls to reduce and/or mitigate impacts;

  • Cyanidation Permit – Required by the IDEQ and is applicable for a facility that processes mineralized material using cyanide as the primary reagent. Integra is proposing to process the gold and silver mineralized material at a heap-leach facility and at a mill with associated TSFs. The regulations apply to both operations and closure and reclamation of any facility;

  • Land Application Permit – In order to apply any treated process wastewater to a designated land area for ultimate disposal, the mining company must obtain a Land Application Permit from IDEQ. The project would need to meet the performance standards for new sources “zero discharge” requirement for net precipitation minus evaporation to ensure no unpermitted discharges. Integra may use underground injection or surface water discharge (as discussed above) rather than land application. Idaho Department of Water Resources (“IDWR”) implements the Underground Injection Control program;

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  • Ground Water Rule – This rule establishes minimum requirements for ground water protection through standards and a set of aquifer protection categories. To implement the rule, Integra would need to request to establish points of compliance outside and down-gradient from the mine area(s). Integra would also establish reasonable upper-tolerance limits for all compliance wells, working directly with IDEQ. These upper-tolerance limits would need to take into account the high naturally occurring background levels for several parameters;

  • Water Rights – There are five decreed water rights associated with the mining activity area. In addition to these established rights, Integra holds three associated permits covering both surface and ground water that can be renewed annually;

  • Stream Channel Alteration Permit – Required by the IDWR for a modification, alteration, or relocation of any stream channel within or below the mean high-water mark. The PFS contemplates relocating one or more unnamed creeks, both temporarily and/or permanently, as part of the overall mine plan;

  • Dam Safety Permit – The IDWR requires a Dam Safety Permit for dams greater than 3.05 m (10 feet) high or for reservoirs exceeding a 50-acre-feet storage capacity. The Application to Construct a Dam includes design plans and specifications for construction of the dam. Mine tailing impoundments greater than or equal to 9.14 meters (30 feet) high are regulated by IDWR in the same manner. Design and Construction Requirements for Mine Tailings Impoundment Structures are described in IDAPA 37.03.05. The PFS contemplates construction of a TSF in an unnamed creek within the Slaughterhouse Gulch drainage and would need authorization;

  • Water and Wastewater Systems – The drinking water system(s) design for the contemplated facilities must be approved prior to use to ensure compliance with the Safe Drinking Water Act;

  • Fuel Storage Facilities – Any proposed fuel storage must also comply with IDEQ design and operating standards, as well as EPA Standards under 40 CFR 112 (may require a Spill Prevention and Countermeasure Plan depending on the size of the facility), Idaho State Fire Marshall, and Owyhee County requirements;

  • Reclamation Plan – All surface mines must submit and obtain approval of a comprehensive reclamation plan (Title 47) for mining activities on patented and state lands. The Reclamation Plan includes detailed operating plans showing pits, mineral stockpiles, overburdened piles, tailing ponds, haul roads, and all related facilities. The Reclamation Plan must also address appropriate BMPs and provide for financial assurance in the amount necessary to reclaim those mining activities. The plan must be approved prior to any surface disturbance. A large portion of the planned Florida Mountain and DeLamar pits are located on patented land;

  • State Historic Preservation Office – The project is located within the DeLamar National Historic District; therefore, approval of a historic/cultural resources assessment by the State Historic Preservation Office would be required; and

  • Others – State requirements would also involve compliance with the Idaho Solid Waste Management Regulations and Standards, transportation safety requirements enforced by the Idaho Public Utilities Commission, and others.

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20.1.4Local County Requirements

There are several other permits and approvals that would apply to the project, if it proceeded to a full-scale mining proposal, including:

  • Conformance with the Owyhee County Comprehensive Plan;

  • Issuance of building permits by the county; and

  • Sewer and water systems approval by Southwest District Health Department, and various other authorizations.

Additionally, an annual authorization by the Owyhee County Road Department for an Owyhee County Road Use Permit for any mining operation is essential. The permit addresses standard operating procedures for the road route to be used, seasonal limits, spill prevention and response planning, HAZWOPER or hazardous materials handling training, convoying, and other requirements.

Integra has not entered into any agreements with local communities; however, there have been discussion with the local communities, principally Silver City and Jordan Valley.

20.1.5Idaho Joint Review Process

The IDL is responsible for implementation of the Idaho Joint Review Process, which was established to coordinate and facilitate the overall mine permitting process in the state. It involves an interagency Memorandum of Understanding (“MOU”) between involved state and federal agencies and addresses a process to achieve pre-analysis coordination in approving/administering exploration permits, interagency agreement on plan completeness, alternatives considered, draft and final permits, bonding during mine plan analysis, and interagency coordination related to compliance, permit changes and reclamation/closure for major mining projects. In Idaho, the Joint Review Process was established as the basis for interagency agreement (state, federal, and local) on all permit review requirements. The focus of the process is concurrent analysis timelines. This would include, for example, in the case of the DeLamar project the NEPA process, IPDES permit, Section 404 permit, Section 401 Certification of these key permits, and the state Cyanidation permit. The Idaho Joint Review Process would play a key role in achieving two primary permitting goals: 1) increased communication and cooperation between the various involved governmental agencies, and 2) reduced conflict, delay, and costs in the permitting process.

20.2EIS/Permitting Timelines and Costs

20.2.1Permitting Timelines

The discussion below assumes that the BLM will be the lead federal agency for NEPA, and that the United States Army Corps of Engineers (“USACE”) will be a cooperating agency. With regard to the likely scope of DeLamar, the following conceptual description was developed as the basis for this permitting analysis:

  • Regulatory – EIS required; BLM Lead Agency; USACE, IDEQ, and IDL are cooperating agencies;

  • Mining- Estimated at 41,000 tonnes per day (heap-leach and mill) mineralized material with a 2.21:1 waste to mineralized material strip ratio;

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  • Processing – Tailing by-product (flotation tailing and concentrate leach tailing); heap leaching of ore is also part of the project. Integra has the option to delay mill construction;

  • Power – It is anticipated that the existing line power would be utilized along with the development of an on-site solar and LNG electricity generation system to meet further power needs;

  • Waste Rock – Some selective placement would be required likely due to potential geochemical reactivity; large volumes would be stored and managed;

  • Water Supply – Available from existing or future water rights;

  • Water Treatment – Required for surface water discharge, injection or land application (similar to the existing system on site);

  • Project Access – Existing Trout Creek Road from Jordan Valley, Oregon with a revised access to the Trout Creek Road within the project area;

  • Manpower – Up to 440 direct and indirect jobs during construction; estimated 340 during operations;

  • Operating Schedule – Mining and processing year-round; and

  • Total Land Disturbance – Over 809 hectares (2,000 acres) of disturbance on public lands, state lands and private lands.

This concept was developed only for the purpose of scaling the project, such that the estimated schedules and costs could be compared with the projects listed earlier.

An EIS/permitting timeline is summarized below in five primary permitting windows.

  • Initial 12 to 24 Months – Selection of third-party EIS Contractor, Start baseline confirmatory studies for surface and ground water, geochemistry, aquatic resources, wildlife, cultural resources, vegetation and soils as well as air quality and wetlands work. This work commenced in 2021. Negotiate an MOU with the BLM for preparing the EIS. Conduct initial internal scoping with “high up” agency and political contacts;

  • Months 20 to 30 – Commence preparation of the Initial Plan of Operations. Concurrently, develop all other permit applications for submittal. The third-party contractor would finalize the EIS work plan and initiate early environmental baseline adequacy determination write-ups for the various resource categories (air, water, socio-economics, etc.);

    • The contractor would also write the alternatives section of the EIS. This is a significant section that must present only reasonable and potentially feasible alternatives as required under NEPA, but also meet the USACE’s alternative analysis requirements. Integra would file its Initial Plan of Operations with the BLM mid-way through this window at around Months 25 to 27, which would trigger the EIS. This is currently anticipated for the first half of 2023;

    • The BLM would publish the Notice of Intent to prepare an EIS during this timeframe;

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  • Integra is completing a feasibility study during this timeline or permitting window. Once this is done, Integra can narrow down the best alternative both from a cost and environmental standpoint. This level of pre-feasibility information would also be crucial to success in obtaining the IPDES water discharge and USACE 404 permits. The USACE 404 Permit would be needed for wetlands disturbance, and any stream diversions or alternation needed for the project;
  • Months 30 to 34 – A preliminary draft EIS would be completed by the BLM (Third-Party Contractor). This document would be for the lead and cooperating agencies and Integra review only. Typically, this review would require about 30 to 60 days. In the initial stages of this period, Integra would file most, if not all, of their permit applications. Some, like the water rights applications, if needed, would have already been submitted to the appropriate agencies. Others, like the USACE 404 Permit, cannot be issued until after the final EIS and ROD by the BLM have been issued;

  • Months 34 to 42 – A draft EIS would be produced for public review. The review period would be about 60 days;

  • At Months 40 to 42, the final EIS would be issued. At this point, the BLM could choose to issue the ROD concurrently or elect to issue it 30 days later. There would be an administrative appeal period involved at this point. For the purposes of this very preliminary assessment, an additional 60 days could be used, pushing the project out to Month 44. The remaining permits would also be issued over this period; and

  • Summary Best Case Schedule – Estimated at 40 to 42 months (3.5 years) with a concurrent baseline data collection program.

20.2.2Most Likely Case EIS Cost Summary

The costs listed below are factored from real costs. This is considered the most realistic estimate given the various permitting uncertainties associated.

  • EIS “Project” – $1.9M (represents all costs for third-party EIS and BLM reimbursement);
  • Support Engineering – $1.0M (does not include feasibility study);
  • Legal – $0.5M;
  • Potential Baseline Study Needs -$7.5M to $9.5M, depending on ground water and geochemistry issues;
  • Permitting “Project” – $3.5M (represents all costs for separate permitting program); and
  • Total Estimate – $14.4M to $16.4M.

20.2.3Integra Permitting Management Strategy

To successfully achieve any such permitting program, estimated costs, and timeline, Integra has designed a seven-point management process that includes the following key points:

  • MOU providing for agency cooperation, accountability, and predictability;
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  • Requirements for quality consultants;

  • Communication plan for the consultants;

  • Baseline studies, adequacy determinations and tracking procedures, EIS completeness;

  • Budget and schedule tracking and cost controls;

  • Goals for environmental enhancement in mine planning and closure; and

  • An informed public affairs process.

20.3Social and Community

The project is located in rural Owyhee County, close to the Oregon border. The closest substantial community is Jordan Valley, in Malheur County Oregon. This community is primarily an agricultural-based economy. However, when the mine previously operated in the 1980s and 1990s many of the employees lived in Jordan Valley.

20.4Dark Skies Compliant Lighting

In order to limit light pollution, the Dark Sky Reserve has created a detailed Lightscape Management Plan. Important considerations include meeting lumen and temperature requirements, shielding fixtures, and reducing glare. The DeLamar site does not fall within a designated Dark Sky area, but the following will be considered in future engineering studies:

●Dark Sky compliance states that any light emitting over 500 initial lumens (40 candela) must be shielded to prevent any light from emitting beyond horizontal. Dark Sky regulations require all lights to be a maximum of 3000k; and

●Lighting fixtures should be installed high and face vertically down. Directional lighting is safer for workers and will drastically reduce the amount of light pollution emitted from a site.

20.5Waste Characterization

Integra’s consultant will be conducting a mine waste characterization program as part of the planning and impact assessment for the project. Geochemical testing of mine waste materials provides a basis for assessment of the potential for metal leaching (“ML”) or acid rock drainage (“ARD”), prediction of contact water quality, and evaluation of options for design, construction, and closure of the mine facilities. This work also supports the next phase of the project’s potential advancement, including environmental assessment and permitting. The characterization effort focuses on the assessment of waste-rock geochemistry, evaluation of tailing material from mineral benefaction processes, and determination of final pit wall geochemistry.

Geochemical characterization is an iterative process and sample collection for the project is being completed in phases. The first phase is complete and involved the collection of samples from core generated during the recent exploration drilling activities. Subsequent phases of the characterization program would focus on improving the spatial representation of the dataset as drill core from the ongoing exploration and geotechnical drilling becomes available.

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20.6Closure and Reclamation Strategy

A comprehensive reclamation and closure plan would be developed for all disturbances and infrastructure associated with the project. Reclamation objective standards established by industry best practices and regulatory requirements for reclamation would be fulfilled. Integra fully understands the importance of monitoring the effect of temperature on fish species that exist in the Jordan Creek drainage, as well as related environmental closure needs required to maintain these resources as part of any overall mine and closure plan. Doing so would involve a “mine for closure strategy” that begins with the end in mind. Integra would seek to develop an economic mine plan and closure/reclamation strategy that integrates habitats and restoration components. The plan would meet all standards of the Clean Water Act. The plan would also mitigate wetland impacts to recreate, enhance or replace productive wetlands and other riparian habitats. It is anticipated that the reclamation and closure of the tailing and heap-leach facilities would consist of fluid management through evaporation or treatment and discharge, covering the facilities growth media, and then revegetating. The estimated reclamation costs for the project, using the Nevada Standardized Reclamation Cost Estimator is approximately $30.8 million. If Integra chooses to delay the construction and operation of the mill facility then the estimated reclamation cost could be approximately $24.8 million.

The goals of this reclamation and closure plan are expected to evolve based on cooperative discussions, public and regulatory input; however, the initial goals include:

  • Protecting water quality;

  • Restricting or eliminating the migration of potential contaminants of concern from all sources based on the proposed mine plan;

  • Restricting or eliminating potential public safety risks associated with the potential decommissioned and reclaimed mine site;

  • Restoring the property, to the extent possible, to the current pre-mining conditions; and

  • Improving the property by incorporating environmental mitigation projects as identified through the NEPA process.

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21.0CAPITAL AND OPERATING COSTS

Capital and operating costs were estimated for the PFS by MDA (mining and infrastructure), M3 Engineering (both heap leaching and milling processing), and Mr. John Welsh, senior principal of Welsh Hagan Associates (leach pads and tailing impoundments). Table 21.1 summarizes the estimated capital costs for the project. The LOM total capital costs are estimated as $589.5 million, including $307.6 million in preproduction and $281.8 million for expansion and sustaining capital. Sustaining capital includes $30.8 million in reclamation costs. Capital costs below are inclusive of sales tax, engineering, procurement, and construction management (“EPCM”) and contingency.

Table 21.1 Capital Cost Summary

Notes:

(1)Capital costs include contingency and EPCM costs;

(2)Mining equipment includes cost of Railveyor.

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(3)Major mining equipment assumes financing by equipment vendor with 10% down and principal payments included under sustaining capital column and interest payments included in operating costs;

(4)Sustaining capital showed in this table includes expansion capital (non-oxide plant) and principal payment of mining equipment leases (see note 3 above);

(5)Working capital is returned in year 17;

(6)Cash deposit = 20% of bonding requirement. Released once reclamation is completed; and

(7)Salvage value for mining equipment and plant.

Table 21.2 shows the estimated LOM operating costs for the project. Operating costs are estimated to be $12.93 per tonne processed for the LOM. This includes mining costs which are estimated to be $1.90 per tonne mined. The total cash cost is estimated to be $923 per ounce of gold equivalent and site-level, all-in sustaining costs are estimated to be $955 per ounce of gold equivalent.

Table 21.2 Operating and Total Cost Summary

Notes:

(1)By-Product costs are shown as US dollars per gold ounces sold with silver as a credit; and

(2)Co-Product costs are shown as US dollars per gold equivalent ounce.

21.1Mining Capital

Mining capital estimates assume owner operations of mining equipment and were based on the equipment and facilities required to achieve the production schedule. Capital costs were based on estimation guides, quotations from equipment vendors and recent costs for similar projects. The mining capital estimate is summarized by year in Table 21.3.

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Table 21.3 Mining Capital Cost by Year

21.1.1Primary Equipment

Primary equipment purchases refer to the purchase of drills, loading equipment, and haul trucks. The total LOM primary equipment cost estimate is $70.8 million which includes:

  • $620,000 for a pioneering drill;

  • $8.5 million for production drills;

  • $2.5 million for a large loader;

  • $11.9 million for hydraulic shovels; and

  • $47.3 million for 136-tonne capacity haul trucks.

Primary equipment costs assume financing with the following terms:

  • Cash deposit upon purchase: 10%

  • Annual interest rate included in operating cost: 5%

  • Financing terms: 5-years

21.1.2Railveyor

The cost estimate to install a Railveyor system to facilitate ore haulage from the edge of the pit to the process facility is estimated at $36.2 million including the following:

  • $9.8 million for train cars;

  • $20.0 million for drive stations;

  • $3.3 million for track;

  • $2.0 million for feeders;

  • $535,000 for installation and commissioning; and

  • $430,000 for control systems.

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21.1.3Support Equipment

Support equipment includes the equipment required to support the primary mining equipment. This includes dozers to manage dumping locations and cleanup of benches for drilling and loading equipment. This also includes road maintenance equipment such as water trucks and graders. The total estimated capital for support equipment is $8.9 million and includes:

  • $4.4 million for dozers of various sizes;

  • $2.5 million for motor graders;

  • $1.4 million for water trucks;

  • $33,000 for in-pit pumps to control runoff water;

  • $557,000 for a 50-ton capacity crane (to be shared between mining and process); and

  • $67,000 for a flatbed truck used for moving maintenance items within the mine.

21.1.4Blasting Equipment

Blasting equipment includes explosives trucks for use in loading blast holes and a skid loader to be used for stemming holes. The cost estimate for blasting equipment is $325,000 which includes $256,000 for one explosives truck and $70,000 for a skid loader.

21.1.5Mine Maintenance Capital

The cost estimate for mine maintenance capital is $1.2 million, which includes one large lubrication/fuel truck at $541,000, two mechanic’s trucks totaling $472,000, and one tire truck totaling $192,000.

21.1.6Other Capital

Other capital includes an assortment of equipment and facilities totaling $2.1 million. This includes:

  • $110,000 for light plants;

  • $200,000 preparation for explosives storage site;

  • $86,000 for ANFO storage bins;

  • $15,000 for powder magazines to store boosters;

  • $8,000 for a cap magazine;

  • $64,000 for mobile radios in equipment and assorted handheld radios;

  • $750,000 for general shop equipment including hoists and other tooling;

  • $105,000 for engineering computers, plotters, and other office equipment;

  • $20,000 for dust suppression storage bladders;

  • $150,000 for surveying equipment and GPS base stations;

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  • $86,000 for fuel island facilities;

  • $400,000 in access roads to each deposit and site preparation; and

  • $150,000 for ambulance and firefighting equipment.

The access roads and shop are described in Section 17.3.

21.1.7Mine Preproduction Costs

Mine preproduction costs are considered as the cost of all mining prior to the start of metal production from the leach pad. For the PFS this is a 12-month period during which there are six months of mining. The mining costs during preproduction total $12.7 million.

21.2Process Capital

The process plant capital costs were developed for the initial phase, oxide and mixed ore processing circuit, as well as for the Stage 2 non-oxide ore processing circuit. The capital costs for each phase are comprised of direct costs and indirect costs. The direct costs were developed from labor, materials, plant equipment, sub-contracts, and construction equipment. Freight is also included with the direct costs. Indirect costs were applied to the direct costs to account for items such as: construction support, EPCM, vendor support during specialty construction and commissioning, spare parts, contingency, owner’s costs, and taxes.

M3 developed the costs for site layout within the process area, the process plant, and several ancillaries. The Stage 1 (Phase 1) process plant (oxide ore) includes the three-stage crushing circuit, ore conveyance and stacking on the heap-leach pads, heap-leach pad solution systems, the Merrill-Crowe gold/silver recovery plant and refinery, and ancillaries (administration building, warehouse, truck shop, maintenance, and metallurgy and assay lab). Stage 1 (Phase 2) process plant (oxide ore) includes agglomeration and additional conveyors for ore stacking on the pad. The Stage 2 process plant (non-oxide ore) includes primary crushed non-oxide ore conveyance, grinding, flotation, concentrate regrind, concentrate leach, concentrate leach counter-current decantation, tailing slurry dewatering, and tailing slurry transport to independent tailing impoundments (rougher flotation tailing and concentrate leach tailing). Welsh-Hagen developed the costs for the heap-leach pads, as well as the tailing impoundments and included freight, sales taxes, EPCM, and contingencies to their estimates.

Indirect costs were then calculated following industry accepted methodologies, including application of contingency based on the completed level of design on a scope or individual work type basis. The agglomerate contingency for the process plant is estimated at 20% of total contracted cost for each phase. Total contracted costs include all process plant direct costs, plus construction support costs, EPCM costs, vendor support costs, spare parts costs, and county taxes. First fills were calculated by M3. Owner’s Costs were defined by Integra and are carried outside of the process plant estimates. Owyhee County sales taxes were included at 6% of plant equipment and material costs. Process plant capital costs estimates were based on the purchase of new equipment.

The total evaluated process plant capital costs are projected to be in the accuracy range of -20% / +25%.

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Stage 1 (Phase 1) process plant capital costs for the oxide and mixed ore types are summarized in Table 21.4. Stage 1 (Phase 2) process plant capital costs for the oxide and mixed ores are summarized in Table 21.5. Stage 2 process plant capital costs for the non-oxide ore are summarized in Table 21.6.

Table 21.4 Stage 1 (Phase 1) Capital Costs, Oxide & Mixed Ore Heap Leach/Merrill Crowe Plant

Category (all costs are in USD 1,000) Labor Plant

Equip.
Material Sub

Contract
Const.

Equip.
Total
General Site (Earthworks) 755 119 376 1,251
Primary Crushing 5,122 4,938 3,501 1,234 14,796
Fine Crushing 10,863 20,204 6,036 2,204 39,307
Conveying/Stacking 1,650 6,525 1,004 59 9,239
Heap Leach Solution Systems 2,466 1,825 3,028 97 7,417
Merrill Crowe 4,304 7,595 2,692 277 14,868
Refinery 628 536 428 69 1,661
Water Systems (In-Plant) 875 2,504 467 45 3,891
Reagents 156 337 74 10 576
Ancillaries 2,522 983 6,323 930 10,757
Freight 3,636 1,894 5,530
Sub-Total Direct Cost 29,343 49,082 25,566 5,301 109,292
Contractor Mobilization 1,948
EPCM 18,355
Vendor Support 1,472
Spare Parts (Capital, Comm.) 2,518
First Fills 750
Owner’s Cost
Taxes (County) 4,147
Contingency 26,717
Sub-Total Indirect Cost 55,907
TOTAL CAPITAL COST 165,198

Table 21.5 Stage 1 (Phase 2) Capital Costs, Oxide & Mixed Ore Heap Leach/Merrill Crowe Plant

Category (all costs are in USD 1,000) Labor Plant

Equip.
Material Sub

Contract
Const.

Equip.
Total
Agglomeration 960 1,755 649 62 3,425
Conveying/Stacking 375 1,375 301 8 2,059
Freight 250 76 326
Sub-Total Direct Cost 1,335 3,380 1,026 70 5,810
Contractor Mobilization 93
EPCM 974
Vendor Support 101
Spare Parts (Capital, Comm.) 186
First Fills
Owner’s Cost
Taxes (County) 245
Contingency 1,433
Sub-Total Indirect Cost 3,032
TOTAL CAPITAL COST 8,842


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Table 21.6 Stage 2 Capital Costs – Non-Oxide Ore Grinding, Flotation, Leach Plant

Category (all costs are in USD 1,000) Labor Plant

Equip.
Material Sub

Contract
Const.

Equip.
Total
General Site 190 190
Primary Crushed Ore Conveyance 1,442 2,178 1,052 161 4,833
Crushed Ore Reclaim 2,241 1,159 1,144 304 4,848
Grinding 9,528 12,545 6,442 820 29,335
Flotation & Concentrate Regrind 5,988 7,081 3,705 573 17,347
Concentrate Leaching 2,496 2,352 1,749 253 6,850
Concentrate CCD 1,341 2,733 772 66 4,912
Flotation Tailing Thickening 4,431 1,669 2,567 993 9,660
Water Systems (In-Plant) 380 978 214 16 1,589
Reagents 1,062 1,965 601 74 3,702
Freight 2,613 1,460 4,072
Sub-Total Direct Cost 29,101 35,271 19,707 3,260 87,339
Contractor Mobilization 1,433
EPCM 14,647
Vendor Support 1,058
Spare Parts (Capital, Comm.) 1,940
First Fills 1,250
Owner’s Cost
Taxes (County) 3,054
Contingency 21,283
Sub-Total Indirect Cost 44,666
TOTAL CAPITAL COST 132,005

21.2.1Freight

Estimates for equipment and material freight costs are based on bulk freight loads and have been estimated at 8% of the equipment and material costs.

21.2.2Construction Support

Mobilization is included as an indirect cost at 5% of total direct field costs for civil contracts and 1.5% for all other total direct field costs for process plant direct costs.

Contractor temporary construction facilities are included with Owner’s Costs. Temporary construction power is included at 0.1% of total direct field cost.

21.2.3EPCM

Engineering is included at 6.0% of total constructed cost (“TCC”) for the process plant scope. Management and accounting are included at 0.75% of TCC. Project services are included at 1.0% of TCC. Project controls are included at 0.75% of TCC. Construction management cost is included at 6.5% of TCC. An EPCM Fee is included at 1.5% of total direct field cost. EPCM construction trailers are expected to be shared with Owner trailers and are included as part of the Owner’s Costs.

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21.2.4Vendor Support

Vendor supervision of specialty construction is included at 1% of plant equipment supply costs. Vendor pre-commissioning and vendor commissioning are each included at 1% of plant equipment supply costs.

21.2.5Spare Parts

Capital spare parts are included at 5.0% of plant equipment supply costs. Commissioning spare parts are included at 0.5% of plant equipment supply costs. Two-year operating spare parts are excluded.

21.2.6Heap-Leach Pad Capital

The estimated cost of the Phase 1 portion of the HLP is $42.3 million and the estimated cost of the Phase 2 portion is $11 million. These costs include contractor mobilizations, earthwork, synthetic liner purchase and installation, freight on materials, sales taxes at 6% of direct cost, EPCM at 4% of direct cost, and 20% for contingency.

21.2.7Tailing Impoundments

The estimated cost of the Phase 1 portion of the Slaughterhouse Gulch TSF is primarily for earthwork to construct the starter embankment and is $24.7 million. Later stages to raise the embankment are estimated to cost $29.8 million in years 5 and 10 of operation. These estimates include contractor costs, sales taxes on materials at 6%, and 20% contingency.

The estimated cost of the CLTSF includes earthwork and purchase and installation of HDPE liner. The estimated cost is $8.1 million. The estimate includes freight and sales taxes at 6% on the synthetic liner, EPCM at 4% of direct cost, and 20% contingency.

21.3Owner and Infrastructure Capital Costs

MDA estimated owner and infrastructure capital costs to be $28.1million for the LOM. This includes costs for power, water, access, security, snow removal equipment, warehouse and offices, light vehicles, and preproduction costs, and excludes preproduction mining costs. The estimated costs are shown in Table 21.7 with further information as follows:

  • Power distribution – $3.5 million based on discussions with Agreeko and micro grid provider;

  • Site water and distribution -$150,000 for two water wells and some water distribution piping;

  • Water treatment plant – $6,100,000 to treat potential contact water discharge;

  • Access road – $8,957,000 for road improvements and access road between deposits (including security and fencing);

  • Security & Fencing – $100,000;

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  • Snow removal equipment – $150,000 for a sanding truck with plow;

  • Light vehicles – $2.1 million (see Section 21.3.1); and

  • Preproduction G&A and process costs – $7.0 million (see Section 21.3.2).

Note: preproduction costs related to mining (pre-stripping) have been included in mining capital (see Section 21.1.7).

Table 21.7 Infrastructure & Owners Capital

21.3.1Light Vehicles

The light vehicle cost is based on the anticipated vehicle need by department for administration, mining general personnel, mine operations personnel, maintenance, and process personnel. The total cost for light vehicles is estimated to be $2,126,000 as shown in Table 21.8. This estimate includes $1,152,000 in initial costs and $974,000 in replacement costs occurring in year 5.

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Table 21.8 Light Vehicle Cost Estimate

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21.3.2Preproduction Owner Costs

Preproduction owner’s costs include G&A and some processing costs during preproduction. Costs related to preproduction mining (pre-stripping) have been included in mining capital (see Section 21.1.7).

The G&A costs include construction management personnel as well as staff for administration, accounting, environmental, and safety and security functions. The total preproduction G&A cost is estimated to be $5.2 million.

Preproduction process costs are estimated to be $1.8 million. This is estimated using fixed and variable costs for tonnage of material sent to the Florida Mountain crusher and leach pad during two months of preproduction. It is assumed that some of this material will be placed on the pad as over-liner material and the remaining of the material will be used for commissioning.

21.4Reclamation Costs and Salvage Value

Reclamation costs were estimated to be approximately $30.8 million at the end of the mine life. The reclamation costs were estimated using the Standardized Reclamation Cost Estimator (version 1.4). The RCE has been developed in accordance with the guidelines created by Nevada Standardized Unit Cost Project, a cooperative effort of the NDEP, the U.S. Department of Interior, Bureau of Land Management, and the Nevada Mining Association.

These costs are estimated based on project specific data. Direct costs of $22.8 million were estimated with an additional $8.0 million in indirect costs.

The reclamation costs were assumed to be secured with a surety bond. A required cash collateral deposit of $6.2 million (20%) was assumed for the surety bond and included in preproduction capital. The surety bond yearly fees were allocated to G&A operating costs.

A credit of 6% for process equipment will be taken as a salvage value at the end of processing. This amounts to $6.9 million in year 17 for process equipment related to leaching, and $5.2 million in year 17 for milling equipment.

A total of $23.7 million in salvage was credited to capital accounts for primary and support mining equipment. This was estimated using the initial cost basis reduced by 20% for initial depreciation, then further reduced based on anticipated life of equipment in terms of hours by equipment type. In the case of the Railveyor system, a 10% salvage value was applied at the end of use. The mining equipment salvage credit was taken in year 17.

21.5Mine Operating Costs

Mine operating costs were estimated using first principals. This was done using estimated hourly costs of equipment and personnel against the anticipated hours of work for each. The equipment hourly costs were estimated for fuel, oil and lubrication, tires, under-carriage, repair and maintenance costs, and special wear items.

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Personnel costs include supervision, operating labor, and maintenance labor. The mine operating costs are summarized by year and category in Table 21.9. Note that while the costs for preproduction are shown in the cost tables below, these costs are capitalized as preproduction costs. Before capitalization of preproduction costs, the LOM mining costs are $756.1 million, or $1.91/tonne mined. This includes $0.04/tonne for lease interest charges.

Table 21.9 Yearly Mine Operating Cost Estimate

21.5.1Mine General Services

Mine general services includes mining supervision along with engineering and geology services. Supervision allows for a mine superintendent, mine general foreman and mine shift foremen. Engineering personnel include a chief engineer along with engineers and surveying crew to support mine planning and operations. Geology is intended to support ore control, geological mapping, and sampling requirements.

Table 21.10 shows the yearly cost estimate for the mine general services.

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Table 21.10 Mine General Services, Engineering and Geology Costs

21.5.2Mine Maintenance

Mine maintenance costs include the cost of personnel for maintenance, supervision, and planning, along with shop support personnel, including light vehicle mechanics, welders, servicemen, tire men, and maintenance labor. The estimated mine maintenance costs are shown in Table 21.11. Note that these costs do not include the maintenance labor directly allocated to the various equipment, which is accounted for in other mining cost categories.

Table 21.11 Yearly Mine Maintenance Costs

21.5.3Drilling

Drilling cost estimates are shown in Table 21.12. The LOM drilling costs are estimated to be $74.6 million or $0.19 per tonne including preproduction. The total LOM cost without capitalized preproduction is $73.5 million.

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Table 21.12 Yearly Drilling Costs

21.5.4Blasting

LOM blasting costs, including preproduction, are shown in Table 21.13. These costs are based on owner operations for blasting and assume ANFO costs of $600/tonne with transportation costs for ANFO at $35/tonne. A blasting accessories cost of $22.00 per hole was included.

The LOM blasting costs are estimated to be $84.0 million or $0.21 per tonne including preproduction. The total LOM cost without capitalized preproduction is $82.9 million.

Table 21.13 Yearly Blasting Costs

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21.5.5Loading

Loading costs are based on operation of two hydraulic shovels with 23-cubic meter (30.0-cubic yards) buckets for all primary production. In addition, a 13-cubic meter (17.0-cubic yard) front-end loader is assumed to be used for stockpile management and re-handling as well as backup for production during shovel maintenance. The yearly loading cost estimate is shown in Table 21.14.

The LOM loading costs are estimated to be $113.7 million or $0.29 per tonne including preproduction. The total LOM cost without capitalized preproduction is $112.1 million.

Table 21.14 Yearly Loading Costs

21.5.6Hauling

Haulage cost was estimated using the truck hour estimates discussed in Section 16.3. The yearly haulage cost estimate is shown in Table 21.15. This includes the use of Railveyor to lower the cost of ore haulage. Of note, the Railveyor cost is reduced for Florida Mountain as it regenerates more power than it consumes.

The LOM haulage costs are estimated to be $301.9 million or $0.76 per tonne including preproduction. The total LOM cost without capitalized preproduction is $297.7 million.

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Table 21.15 Yearly Haulage Costs

21.5.7Mine Support

Yearly mine support cost estimates are shown in Table 21.16 including preproduction costs. These costs assume the hourly costs for required support equipment and personnel as discussed in Sections 16.3 and 16.4 respectively.

The LOM support costs are estimated to be $85.0 million or $0.21 per tonne including preproduction. The total LOM cost without capitalized preproduction is $82.8 million.

Table 21.16 Yearly Mine Support Costs

21.6Process Operating Cost Summary

Process operating costs were developed by M3 from first principles. Labor costs were estimated using project specific staffing, salary and wage, and benefit requirements. Unit consumptions of materials, supplies, power, and delivered supply costs were also estimated. The consumptions of cyanide and lime vary between the ore domains. These are applied in the financial model, outside of the static cost per tonne, to allow for the variations in reagent consumptions to be allocated according to the source of the ore. Table 21.17 lists the average operating costs per tonne at full plant capacity for the oxide and mixed heap-leach plant and for the non-oxide plant. As aforementioned, the individual costs per tonne vary in the financial model to apply the cyanide and lime consumptions per domain. Furthermore, the process operating overall costs per year are also dependent on quantity of each ore type as well as the total power consumptions.

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Table 21.17 Operating Costs Ratios for Deposit/Ore Types

Deposit DeLamar DeLamar Florida

Mtn
Florida

Mtn
DeLamar Florida

Mtn
Ore Type Oxide Mixed Oxide Mixed Non-Oxide Non-Oxide
Process Plant Capacity (tpd) 35,000 35,000 35,000 35,000 6,000 6,000
Operating Labor ($/t) $0.27 $0.27 $0.27 $0.27 $1.65 $1.65
Power ($/t) $0.35 $0.35 $0.35 $0.35 $1.74 $1.74
Consumables ($/t) $1.82 $2.51 $1.43 $2.71 $7.48 $5.29
Maintenance ($/t) $0.27 $0.27 $0.27 $0.27 $1.38 $1.38
Supplies & Services ($/t) $0.08 $0.08 $0.08 $0.08 $0.55 $0.55
Total Process Plant ($/t) $2.78 $3.47 $2.39 $3.67 $12.80 $10.60

Operating costs were estimated based on 4th quarter 2021 US dollars and are presented with no added contingency based upon the design and operating criteria present in this Technical Report. Operating costs are considered to have an accuracy of +/- 20%.

The process operating costs presented are based upon the ownership of all process production equipment and site facilities. The owner will employ and direct all operating maintenance and support personnel for all site activities.

Operating costs estimates have been based upon information obtained from the following sources:

  • Project metallurgical test work and process engineering;

  • Development of a detailed equipment list and demand/consumption calculations;

  • M3 in-house data for reagent pricing; and

  • Experience with other similar operations.

Where specific data does not exist, cost allowances have been based upon consumption and operating requirements from other similar properties for which reliable data exist.

21.6.1Power

Power usage for the process and process facilities was derived from estimated connected loads assigned to powered equipment from the mechanical equipment list. Equipment power demands under normal operation were assigned and coupled with estimated on-stream times to determine the average energy usage and cost. Power requirements for the heap-leach and Merrill-Crowe operation are presented in Table 21.18. Power requirements for the non-oxide mill operation are presented in Table 21.19.

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Table 21.18Power Summary for Heap Leach & Merrill Crowe Facility

Area Description Connected

Power

(kW)
Demand

Power

(kW)
Consumed

Power

(kW)
Annual

Consumption

(kWh)
Primary Crushing 1,607 1,406 865 7,469,439
Fine Crushing 5,798 5,146 3,520 30,413,698
Agglomeration 436 389 266 2,302,337
Conveying/Stacking 3,158 2,691 1,684 14,552,257
Heap Leach Solution Systems 6,867 1,787 1,730 14,943,253
Merrill Crowe 1,795 1,466 1,436 12,410,402
Refinery 368 198 194 1,673,799
Water Systems 795 406 398 3,437,362
Reagents 25 8 8 71,288
Compressed Air 57 26 26 221,710
Total Process Plant 20,907 13,524 10,127 87,495,544

Table 21.19 Power Summary for Non-Oxide Mill Facility

Area Description Connected

Power

(kW)
Demand

Power

(kW)
Consumed

Power

(kW)
Annual

Consumption

(kWh)
Primary Crushed Ore Conveyance 1,141 171 103 891,888
Crushed Ore Reclaim 204 168 134 1,159,514
Grinding 7,094 6,324 5,518 47,673,955
Flotation & Concentrate Regrind 2,309 2,034 1,871 16,167,642
Concentrate Leaching 259 210 191 1,652,903
Concentrate CCD 419 305 277 2,391,342
Flotation Tailing Thickening 498 258 235 2,028,972
Water Systems 297 252 232 2,002,910
Reagents 252 146 135 1,163,454
Total Process Plant 12,473 9,869 8,696 75,132,581

21.6.2Consumable Items

Operating supplies have been estimated based upon unit costs and consumption rates projected by metallurgical tests, which are detailed in Section 13 of this report. Freight costs are included in all operating supply and reagent estimates. Reagent consumptions have been derived from test work and from design criteria considerations. Other consumable items have been estimated by M3 based on experience with other similar operations.

21.6.3Maintenance

Annual maintenance costs have been included for the process facilities. The maintenance costs are estimated from the capital cost of the plant equipment at an allowance of 5% for parts repair or replacement. Maintenance labor is also included. The maintenance labor for the heap-leach and Merrill-Crowe operation includes one maintenance supervisor, four mechanics, and two electricians. These personnel are included as part of the overall process personnel quantity. The maintenance labor for the non-oxide mill operation includes one maintenance supervisor, eight mechanics, and two electricians. These personnel are included as part of the overall process personnel quantity. An allowance for outside repairs is also included at 10% of the maintenance parts allowance.

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21.6.4Supplies and Services

Estimates for supplies and services have been included for items such as lubricants, third-party services for the process plant, safety items, and minor supplies and tools outside of maintenance.

21.6.5Application of Operating Costs

The process costs were developed by M3 based on the nominal processing rates. Mr. Dyer created the production schedule using the nominal processing rates as production targets, but due to mining in different areas, the rates were not always met. Thus, the process costs needed to be adjusted for periods where material will not be available to maximize processing rates. Mr. Dyer worked with M3 to determine which cost components to consider as fixed costs and which components were to be used as variable costs.

The fixed costs are those costs that remain relatively constant by time rather than changing based on the amount of tonnage being processed per tonne. The fixed costs considered labor, maintenance, and supplies and services as fixed costs which were applied monthly. These costs were developed using the $/t cost calculated by M3 and multiplying them against the nominal tonnage to be processed in each monthly period.

Later in the mine schedule, starting around year 9, the amount of leach material reporting to the plant will be substantially reduced. This would mean that the labor, maintenance, and supplies and services will also be significantly reduced, and though they will not become non-existent, they will have some cost to them. For this reason, once the amount of leach material reporting to the heap-leach plant becomes reduced to below 50% of the nominal rate (triggered in year 9), then the fixed costs will be reduced by 90%.

The variable costs are those that vary based on the tonnage processed. Costs for power and consumables were considered variable costs and were applied using a rate per tonne processed as provided by M3.

The modified process operating costs used in the economic model are summarized by year in Table 21.20. The resulting costs exceed the overall cost per tonne estimated by M3 slightly due to the additional application of the fixed costs.

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Table 21.20 Modified Process Operating Costs

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Mr. Dyer also modified M3’s power costs. Based on studies for power costing, M3 was given an original cost of power of $0.05/kWh. Prior to completion of the PFS, the cost of power was revised to 0.065/kWh. This would represent the realized weighted power cost based on LNG and solar generation. The result is an increase of $0.101 cost per tonne of heap leach ore and $0.521 cost per tonne of mill ore due to the revised power cost.

21.7G&A Costs

G&A costs were estimated based on personnel requirements for administrative, accounting, safety and security, and environmental departments to support mining and processing activities. Costs are also included for legal, land, permit bonding, power, etc. Table 21.21 shows the yearly G&A cost estimate.

Note that preproduction costs are capitalized as part of the owner’s costs (see Section 21.3.2). The resulting LOM G&A cost is $110.8 million. The cost after capitalization of preproduction costs is $105.7 million.

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Table 21.21 Yearly G&A Costs

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22.0ECONOMIC ANALYSIS

MDA has prepared this PFS for the DeLamar mining project, which includes operations at both DeLamar and Florida Mountain. The economic analysis uses principal assumptions which include the metal recoveries as discussed in Section 13, processing methods discussed in Section 17, metal prices discussed in Section 19, and the capital and operating costs from Section 21.

A summary of the PFS economic analysis is shown in Table 22.2. Some economic highlights include:

  • Initial construction period is anticipated to be 18 months;

  • After-tax net present value (“NPV”) (5%) of $407.8 million with a 27% after-tax internal rate of return (“IRR”) using $1,700 and $21.50 per ounce gold and silver prices, respectively;

  • After-tax payback period of 3.34 years;

  • Year 1 to 8 gold equivalent average production of 163,000 ounces (average 121,000 oz Au/year and 3,312,000 oz Ag/year);

  • Year 1 to 16 gold equivalent average production of 110,000 ounces (average 71,000 oz Au/year and 3,085,000 oz Ag/year).

  • After-tax LOM cumulative cash flow of $689.3 million; and

  • Average annual after-tax free cash flow of $59.8 million during production.

Figure 22.1 shows the annual operating after-tax cash flow.

Figure 22.1 Annual Operating After-Tax Cash Flow

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22.1Mining Physicals

The cash-flow model uses the mining and process production schedule physicals summarized in Table 22.1. and as discussed in Section 16.2. Gold and silver ounce production is also shown in Table 22.1 based on the metal production model provided by Mr. Botz.

The payable metal was compiled and shown in the revenue section of the cash-flow mode in Table 22.2. This assumes a 99.5% payable factor applied to the metal production in Table 22.1. The payable metal production is also shown by process method in Table 22.2 and Figure 22.3 shows the metal profile by gold and silver.

Of note: The LOM average recovery for Florida Mountain leach material is 76% and 47% for gold and silver, respectively. The LOM average DeLamar recovery for gold and silver is 66% and 32%, respectively. For mill ore, the LOM averages are 83% and 72% for Florida Mountain gold and silver, respectively. DeLamar non-oxide LOM average mill recoveries are 37% and 75% for gold and silver respectively. These recovery numbers are prior to the application of the 99.5% payable metal factor. With the payable factor, the overall LOM payable recoveries are 64% and 54% for gold and silver, respectively.

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Table 22.1 Yearly Mine & Process Physicals

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Figure 22.2 Gold Equivalent Production Profile by Process Method

Figure 22.3 Gold Equivalent Profile by Process Metals

22.2Pre-Tax Cash Flow

The pre-tax cash-flow model is shown in Table 22.2. This is based on the mining physicals shown in Table 22.1 along with the applications of metal prices discussed in Section 19.0 and the operating and capital costs discussed in Section 21.0 The revenues are based on $1,700 and $21.50 per ounce gold and silver prices, respectively. Transportation and refining costs are assumed to be $5.00 per ounce of gold and $0.50 per ounce Ag of silver produced. Royalties have been applied as NSR royalties described in