EX-99.1 2 exhibit99-1.htm TECHNICAL REPORT UPDATED FEASIBILITY STUDY, PIMA COUNTY, ARIZONA Augusta Resource Corporation - Exhibit 99.1 - Filed by newsfilecorp.com




ROSEMONT COPPER PROJECT
FORM 43-101F1 TECHNICAL REPORT

DATE AND SIGNATURES PAGE

The Qualified Persons contributing to this report are noted below. The Certificates and Consent forms of the Qualified Persons are located in Appendix A, Certificate of Qualified Persons (QP) and Consent of Authors.

  • Dr. Conrad E. Huss, P.E., Ph.D.; Senior Vice President and Chairman of the Board with M3 Engineering & Technology Corporation (M3) in Tucson, Arizona; principal author of this technical report and responsible for Sections 1 through 6, Sections 18 through 22, and Sections 24 through 27 of this report.

  • Susan C. Bird, M.Sc., P. Eng.; Senior Associate with Moose Mountain Technical Services (MMTS) in Cranbrook, British Columbia, Canada; responsible for Sections 7 through 12, Section 14, and Section 23 of this report.

  • Thomas L. Drielick, P.E.; Senior Vice President with M3 Engineering & Technology Corporation (M3) in Tucson, Arizona; responsible for Section 13 and Section 17 of this report.

  • Robert H. Fong, P. Eng.; Principal Mining Engineer with Moose Mountain Technical Services (MMTS) in Cranbrook, British Columbia, Canada; responsible for Section 15 and Sections 16.1 to 16.6 of this report.

  • John I. Ajie, P.E.; Vice President of Engineering Civil Construction & Mining Group with URS Energy and Construction in Denver, Colorado; responsible for the mining capital and operating costs in Section 21 of this report.

This Technical Report is current as of August 28, 2012. The effective date of the report is the issue date of the press release, which was July 24, 2012.

(Signed) (Sealed) “Conrad E. Huss”   August 28, 2012
Conrad E. Huss, P.E., Ph.D.   Date
     
(Signed) “Susan C. Bird”   August 28, 2012
Susan C. Bird, M.Sc., P. Eng.   Date
     
(Signed) “Thomas L. Drielick”   August 28, 2012
Thomas L. Drielick, P.E.   Date
     
(Signed) “Robert H. Fong”   August 28, 2012
Robert H. Fong, P. Eng.   Date
     
(Signed) (Sealed) “John I. Ajie”   August 28, 2012
John I. Ajie, P.E.   Date

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TABLE OF CONTENTS

SECTION PAGE

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 FORM 43-101F1 TECHNICAL REPORT

DATE AND SIGNATURES PAGE I
         
TABLE OF CONTENTS II
         
LIST OF FIGURES AND ILLUSTRATIONS IX
         
LIST OF TABLES XII
         
1 SUMMARY 1
         
  1.1 PROPERTY 1
         
  1.2 LOCATION 1
         
  1.3 OWNERSHIP 1
         
  1.4 GEOLOGY AND MINERALIZATION 2
         
  1.5 EXPLORATION AND SAMPLING 2
         
  1.6 MINERAL RESOURCE 2
         
  1.7 MINE RESERVES & MINE PLAN 5
         
  1.8 MINERAL PROCESSING & METALLURGICAL TESTING 6
         
  1.9 RECOVERY METHODS 8
         
  1.10 ENVIRONMENTAL 8
         
  1.11 ECONOMIC ANALYSIS 8
         
    1.11.1 Operating Costs 8
    1.11.2 Initial Capital Cost 9
    1.11.3 Financial Analysis 9
         
  1.12 CONCLUSIONS & RECOMMENDATIONS 10
         
2 INTRODUCTION 11
         
  2.1 GENERAL 11
         
  2.2 PURPOSE OF REPORT 11
         
  2.3 SOURCES OF INFORMATION 11
         
  2.4 CONSULTANTS AND QUALIFIED PERSONS 11
         
  2.5 DEFINITION OF TERMS USED IN THIS REPORT 13
         
3 RELIANCE ON OTHER EXPERTS 16
         
4 PROPERTY DESCRIPTION AND LOCATION 18

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  4.1 LOCATION 18
         
  4.2 LAND TENURE 18
         
5 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY 21
         
  5.1 ACCESSIBILITY 21
         
  5.2 CLIMATE 21
         
  5.3 LOCAL RESOURCES 22
         
  5.4 INFRASTRUCTURE 22
         
  5.5 PHYSIOGRAPHY 23
     
6 HISTORY 25
         
7 GEOLOGICAL SETTING AND MINERALIZATION 28
         
8 DEPOSIT TYPES 35
         
9 EXPLORATION 36
         
10 DRILLING 37
         
  10.1 BANNER MINING COMPANY DRILLING 38
         
  10.2 THE ANACONDA COMPANY DRILLING 38
         
  10.3 ASARCO MINING COMPANY DRILLING 39
         
  10.4 AUGUSTA DRILLING 39
         
11 SAMPLE PREPARATION, ANALYSES AND SECURITY 41
         
  11.1 BANNER, ANACONDA AND ANAMAX SAMPLING AND ANALYSES 41
         
  11.2 ASARCO SAMPLING AND ANALYSES 42
         
  11.3 AUGUSTA SAMPLING AND ANALYSES 42
         
  11.3.1 Augusta Core 42
11.3.2 Banner, Anaconda, Anamax, and ASARCO Core Resampling and Analysis 44
         
  11.4 SAMPLE HANDLING AND SECURITY 45
         
  11.5 QUALITY ASSURANCE AND QUALITY CONTROL 45
         
  11.5.1 Historic Protocols 45
  11.5.2 Augusta Protocols 46
  11.5.3 External Augusta Standard Reference Materials 46
  11.5.4 Internal Skyline Laboratory Standard Reference Materials 47
  11.5.5 Blank Samples 50
  11.5.6 Check Assays (Pulp Rerun Analysis) 50
         
  11.6 QA/QC SUMMARY 53

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12 DATA VERIFICATION 54
         
  12.1 TEN HOLE RESAMPLING PROGRAM 54
         
  12.2 ADJACENT (METALLURGICAL) COMPARISON HOLES 55
         
  12.3 DRILL HOLE DATA ENTRY VALIDATION 57
         
13 MINERAL PROCESSING AND METALLURGICAL TESTING 58
         
14 MINERAL RESOURCE ESTIMATES 64
         
  14.1 DRILL HOLE DATABASE 64
         
  14.1.1 Assay Statistics and Domain Definition 65
  14.1.2 Compositing of Drill Hole Data 68
         
  14.2 VARIOGRAPHY 72
         
  14.3 BLOCK MODEL INTERPOLATION AND RESOURCE CLASSIFICATION 74
         
  14.3.1 Resource Classification 75
         
  14.4 BLOCK MODEL VALIDATION 76
         
  14.4.1 Comparison of Mean Grades 76
  14.4.2 Grade-Tonnage Curves 76
  14.4.3 Swath Plots 82
  14.4.4 Visual Validation 85
         
  14.5 IN SITU MINERAL RESOURCE ESTIMATE 91
         
  14.6 ADDITIONAL MINERAL RESOURCE POTENTIAL 95
         
15 MINERAL RESERVE ESTIMATES 96
         
  15.1 MINING OVERVIEW 96
         
  15.2 GEOTECHNICAL RECOMMENDATIONS 96
         
  15.3 PIT OPTIMIZATION 99
         
  15.3.1 Metallurgical Recoveries 100
  15.3.2 Economic Parameters 100
  15.3.3 Slope Angles 102
  15.3.4 Lerchs-Grossman Analyses 103
         
  15.4 MINING PHASE DESIGNS 110
         
  15.4.1 Pit Design Parameters 110
  15.4.2 Mining Phases and Ultimate Pit 111
         
  15.5 MINERAL RESERVES 118
         
  15.5.1 Ore Definition Parameters 118
  15.5.2 Material Densities 119
  15.5.3 Dilution 119
  15.5.4 Mineral Reserve Estimates 120

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16 MINING METHODS 124
         
  16.1 PRODUCTION SCHEDULING CRITERIA 124
         
  16.2 MILL FEED CUT-OFF GRADE STRATEGY 125
         
  16.3 OVERBURDEN STRIPPING REQUIREMENTS 126
         
  16.4 WASTE ROCK AND TAILINGS STORAGE 127
         
    16.4.1 Waste Rock Storage Design Criteria 127
         
  16.5 MINE PLAN 128
         
    16.5.1 Mine Preproduction Summary 156
    16.5.2 Mine Production 158
         
  16.6 MINE PRODUCTION SCHEDULE 163
         
    16.6.1 Waste Rock Storage Destinations and Quantities 165
         
  16.7 MINE EQUIPMENT SELECTION 166
         
  16.8 EQUIPMENT OPERATING PARAMETER 168
         
  16.9 DRILLING AND BLASTING 170
         
  16.10 LOADING 170
         
  16.11 HAULING 171
         
  16.12 SUPPORT EQUIPMENT 173
         
  16.13 EQUIPMENT OPERATING HOUR REQUIREMENTS 173
         
  16.14 MINE PERSONAL REQUIREMENTS 175
         
17 RECOVERY METHODS 177
         
18 PROJECT INFRASTRUCTURE 179
         
  18.1 ACCESS ROADS AND PLANT ROADS 179
         
    18.1.1 Main (East) Access Road 179
    18.1.2 Secondary (West) Access Road 180
    18.1.3 In Plant Roads 180
    18.1.4 Haul Roads 180
    18.1.5 Perimeter Road 180
         
  18.2 POWER SUPPLY AND DISTRIBUTION 181
         
    18.2.1 Transmission Line Route 181
    18.2.2 Electrical Power Supply Description 181
         
  18.3 WATER SUPPLY AND DISTRIBUTION 182
         
    18.3.1 Water Supply 182
    18.3.2 Delivery System 183
    18.3.3 Plant Water Distribution 184
    18.3.4 Recharge Plan 185

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  18.4 WASTE MANAGEMENT 185
         
  18.4.1 Hazardous Waste 186
    18.4.2 Non-hazardous Waste 187
         
  18.5 TAILINGS MANAGEMENT 187
         
  18.5.1 General 187
  18.5.2 Location and Design Criteria 187
  18.5.3 Site Conditions and Geology 188
  18.5.4 Tailings Properties 189
  18.5.5 Dry Stack Stability 189
  18.5.6 Hydrologic Modeling 190
  18.5.7 Dry Stack Operations 191
  18.5.8 Surface Water Control 191
         
  18.6 TRANSPORTATION AND SHIPPING 192
         
  18.6.1 Copper and Moly Concentrates 192
  18.6.2 Pebble Lime 193
  18.6.3 SAG and Ball Mill Grinding Balls 193
  18.6.4 Diesel Fuel 193
  18.6.5 Ammonium Nitrate 193
  18.6.6 Miscellaneous Consumables 194
  18.6.7 Miscellaneous Fuels and Lubricants 194
  18.6.8 Employees 194
  18.6.9 Safety Evaluation 195
         
  18.7 COMMUNICATIONS 195
         
  18.8 GEOTECHNICAL STUDIES 196
         
  18.8.1 Geotechnical Study 196
  18.8.2 Geochemical Study 200
  18.8.3 Geologic Hazards Assessment 201
         
19 MARKET STUDIES AND CONTRACTS 207
         
  19.1 COPPER CONCENTRATES SALES 207
         
  19.2 MOLYBDENUM CONCENTRATES 208
         
  19.3 COPPER CONCENTRATE TRANSPORTATION 208
         
20 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT 212
         
  20.1 POLICY, LEGAL, AND REGULATORY FRAMEWORK 212
         
  20.1.1 National Environmental Policy Act (NEPA) 212
20.1.2 National Historic Preservation Act (NHPA) / Cultural Resources Clearances 213
20.1.3 Endangered Species Act (ESA) and Other Biological Requirements 213

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  20.1.4 Water Permitting 214
  20.1.5 Air Permitting 215
  20.1.6 Arizona Native Plant Law 216
  20.1.7 Pima County Conservation Lands System 216
         
  20.2 BASELINE DATA 218
         
  20.3 ENVIRONMENTAL IMPACTS 218
         
  20.3.1 Climate 218
  20.3.2 Air Quality 218
  20.3.3 Soils 218
  20.3.4 Surface Water 219
  20.3.5 Groundwater 220
  20.3.6 Vegetation and Wildlife 221
  20.3.7 Socio-Economics 222
  20.3.8 Visual Resources 222
  20.3.9 Other Resources (Noise, Light, Recreation) 223
         
  20.4 MITIGATION PLAN 224
         
21 CAPITAL AND OPERATING COSTS 226
         
  21.1 INITIAL CAPITAL COST 226
         
  21.2 SUSTAINING CAPITAL COST 228
         
  21.3 OPERATING COSTS 230
         
  21.3.1 Mine Operating Cost 231
  21.3.2 Mill Operations 231
  21.3.3 Supporting Facilities 231
         
22 ECONOMIC ANALYSIS 233
         
  22.1 INTRODUCTION 233
         
  22.2 MINE PRODUCTION STATISTICS 233
         
  22.3 PLANT PRODUCTION STATISTICS 233
         
  22.4 SMELTER RETURN FACTORS 234
         
  22.5 CAPITAL EXPENDITURES 235
         
  22.5.1 Initial Capital 235
  22.5.2 Sustaining Capital 235
  22.5.3 Working Capital 235
  22.5.4 Salvage Value 235
         
  22.6 REVENUE 235
         
  22.7 CASH COPPER UNIT COST NET OF BY-PRODUCT CREDITS 238
         
  22.7.1 Pre-production Mining Cost 239
  22.7.2 Fees and Royalties 239

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  22.7.3 Depreciation 239
  22.7.4 Depletion 239
  22.7.5 Income Taxes 239
         
  22.8 PROJECT FINANCING 240
         
  22.8.1 Net Income After Tax 240
22.8.2 Net Present Value, Internal Rate of Return and Sensitivity Analysis 240
         
23 ADJACENT PROPERTIES 243
         
24 OTHER RELEVANT DATA AND INFORMATION 244
         
25 INTERPRETATION AND CONCLUSIONS 245
         
  25.1 CONCLUSIONS 245
         
  25.2 RISKS 245
         
26 RECOMMENDATIONS 247
         
27 REFERENCES 248
         
APPENDIX A: FEASIBILITY STUDY CONTRIBUTORS AND PROFESSIONAL QUALIFICATIONS 251
         
APPENDIX B: UNPATENTED CLAIMS LIST 266
         
APPENDIX C: PATENTED CLAIMS & FEE LAND LIST 271
         
APPENDIX D: LG SET 1 CASES 276
         
APPENDIX E: RESERVES REPORT 278
         
APPENDIX F: BENCHES MINED BY PERIOD 326

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LIST OF FIGURES AND ILLUSTRATIONS

FIGURE DESCRIPTION PAGE
     
Figure 4-1: Property Location of Rosemont Project 18
   
Figure 4-2: Rosemont Property Ownership 20
   
Figure 7-1: Regional Geology of the Rosemont Property 29
   
Figure 7-2: Rosemont District Stratigraphic Column 30
   
Figure 7-3: Rosemont Deposit Geology – 4,000 Foot Level Plan 31
   
Figure 7-4: Rosemont Deposit Geologic Section – 11,554,825 E 32
   
Figure 10-1: Drill Hole Locations in Rosemont Deposit 37
   
Figure 11-1: QA/QC Standard and Blank Performance 48
   
Figure 11-2: Check Assay Performance 52
   
Figure 14-1: CPP Plots of each Metal for All Domains – Oxide Zone 69
   
Figure 14-2: CPP Plots of each Metal for All Domains – Mixed Zone 70
   
Figure 14-3: CPP Plots of TCu for Sulfide Domains requiring Capping 70
   
Figure 14-4: CPP Plots of Mo for Sulfide Domains requiring Capping 71
   
Figure 14-5: CPP Plots of Ag for Sulfide Domains Requiring Capping 71
   
Figure 14-6: Major Axis Correlogram – TCu 73
   
Figure 14-7: Minor Axis Correlogram - TCu 73
   
Figure 14-8: Grade-Tonnage Curve for TCu-Sulfide Zone 77
   
Figure 14-9: Grade-Tonnage Curve for TCu-Mixed Zone 78
   
Figure 14-10: Grade-Tonnage Curve for TCu-Oxide Zone 78
   
Figure 14-11: Grade Tonnage Curve for Mo-Sulfide Zone 79
   
Figure 14-12: Grade-Tonnage Curve for Mo-Mixed Zone 79
   
Figure 14-13: Grade-Tonnage Curve for Mo-Oxide Zone 80
   
Figure 14-14: Grade-Tonnage Curve for Ag-Sulfide Zone 80
   
Figure 14-15: Grade-Tonnage Curve for Ag-Mixed Zone 81
   
Figure 14-16: Grade-Tonnage Curve for Ag-Oxide Zone 81
   
Figure 14-17: Swath Plots for TCu-Sulfide Zone 83
   
Figure 14-18: Swath Plots for Mo – Sulfide Zone 84
   
Figure 14-19: Swath Plots for Ag – Sulfide Zone 85
   
Figure 14-20: E-W Section at 11,554,825N of OK Model and Assays – TCu Grades 86
   
Figure 14-21: E-W Section at 11,554,825N of OK Model and Assays – Mo Grades 87

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Figure 14-22: E-W Section at 11,554,825N of OK Model and Assays – Ag Grades 88
   
Figure 14-23: Plan at 4000’ of OK Model and Assays – TCu Grades 89
   
Figure 14-24: Plan at 4000’ of OK Model and Assays – Mo Grades 90
   
Figure 14-25: Plan at 4000’ of OK Model and Assays – Ag Grades 91
   
Figure 15-1: Pit Slope Design Sectors and Maximum Slope Angles 97
   
Figure 15-2: Arkose Interramp Slope Angles 99
   
Figure 15-3: Plan View Contours of Selected Lerchs-Grossman Pit Shell 105
   
Figure 15-4: East West Section View of Selected Lerchs-Grossman Pit Shell 106
   
Figure 15-5: Sensitivity Analysis on Metal Prices 109
   
Figure 15-6: Sensitivity Analysis on Mine Operating Costs 110
   
Figure 15-7: Plan View of Mining Pit Phase 1 112
   
Figure 15-8: Plan View of Mining Pit Phase 2 113
   
Figure 15-9: Plan View of Mining Pit Phase 3 114
   
Figure 15-10: Plan View of Mining Pit Phase 4 115
   
Figure 15-11: Plan View of Mining Pit Phase 5 116
   
Figure 15-12: Plan View of Mining Pit Phase 6 117
   
Figure 15-13: Plan View of Mining Pit Phase 7 (Ultimate Pit) 118
   
Figure 16-1: Mine Plan Site Layout 129
   
Figure 16-2: Mine Plan end of Period Pre-Production Q1 130
   
Figure 16-3: Mine Plan End of Period Pre-Production Q2 131
   
Figure 16-4: Mine Plan End of Period Pre-Production Q3 132
   
Figure 16-5: Mine Plan End of Period Pre-Production Q4 133
   
Figure 16-6: Mine Plan End of Period Pre-Production Q5 134
   
Figure 16-7: Mine Plan End of Period Pre-Production Q6 135
   
Figure 16-8: Mine Plan End of Period Pre-Production Q7 136
   
Figure 16-9: Mine Plan End of Period Year 01Q1 137
   
Figure 16-10: Mine Plan End of Period Year 01Q2 138
   
Figure 16-11: Mine Plan End of Period Year 01Q3 139
   
Figure 16-12: Mine Plan End of Period Year 01Q4 140
   
Figure 16-13: Mine Plan End of Period Year 02 Q1 141
   
Figure 16-14: Mine Plan End of Period Year 02Q2 142
   
Figure 16-15: Mine Plan End of Period Year 02Q3 143

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Figure 16-16: Mine Plan End of Period Year 02Q4 144
   
Figure 16-17: Mine Plan End of Period Year 03 145
   
Figure 16-18: Mine Plan End of Period Year 04 146
   
Figure 16-19: Mine Plan End Period Year 05 147
   
Figure 16-20: Mine Plan End of Period Year 06 148
   
Figure 16-21: Mine Plan End of Period Year 07 149
   
Figure 16-22: Mine Plan End of Period Year 08 150
   
Figure 16-23: Mine Plan End of Period Year 09 151
   
Figure 16-24: Mine Plan End of Period Year 10 152
   
Figure 16-25: Mine Plan End of Period Year 15 153
   
Figure 16-26: Mine Plan End of Period Year 19 154
   
Figure 16-27: Mine Plan End of Period Year 21.3 155
   
Figure 17-1: Overall Process Flowsheet 178
   
Figure 18-1: Roads 203
   
Figure 18-2: Transmission Line 204
   
Figure 18-3: Waterline 205
   
Figure 18-4: Typical Easement 206

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LIST OF TABLES

TABLE DESCRIPTION PAGE
     
Table 1-1: Base Case Recoveries, Metal Prices and Resulting Net Smelter Prices 3
   
Table 1-2: Base Case Mineral Resource by Classification and Zone 4
   
Table 1-3: Combined Proven and Probable Mineral Reserves by Phase 6
   
Table 1-4: Estimated Metal Recovery by Year of Production 7
   
Table 1-5: Summary of Average Life of Mine Operating Costs 9
   
Table 1-6: Base Case and Historical Metals Prices 10
   
Table 1-7: Long Term Metals Prices 10
   
Table 1-8: Financial Indicators (After Tax) 10
   
Table 10-1: Rosemont Deposit Drilling Summary 37
   
Table 11-1: Rosemont and Skyline Standards Qa/Qc Results 49
   
Table 11-2: Check Assay Performance Summary 53
   
Table 12-1: Ten Hole Resampling Program Summary 54
   
Table 12-2: Adjacent Drill Hole Comparison Summary 56
   
Table 13-1: Grinding Mill Sizing Parameters 59
   
Table 13-2: Molybdenite Flotation 60
   
Table 13-3: Lithology of Composite Samples 61
   
Table 13-4: Estimated Metal Recovery by Year of Production 62
   
Table 13-5: Estimated Reagent Consumption Rates 63
   
Table 14-1: Domain Definition based on Lithology 65
   
Table 14-2: Assay Statistics of TCu for Domains 8 through 10 – Oxides 66
   
Table 14-3: Assay Statistics of TCu for Domains 1 through 9 – Sulfides 66
   
Table 14-4: Assay Statistics of TCu for Domains 10 through 21 – Sulfides 66
   
Table 14-5: Assay Statistics of TCu for Domains 1 through 21 – Mixed 67
   
Table 14-6: Assay Statistics of Mo for Domains 8 through 10 – Oxides 67
   
Table 14-7: Assay Statistics of Mo for Domains 1 through 9 – Sulfides 67
   
Table 14-8: Assay Statistics of Mo for Domains 10 through 21 – Sulfides 68
   
Table 14-9: Assay Statistics of Mo for Domains 1 through 21 – Mixed 68
   
Table 14-10: Capping Value of Composites 69
   
Table 14-11: Variogram Parameters 72
   
Table 14-12: Block Model Dimensions 74

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Table 14-13: Specific Gravity by Lithology 74
   
Table 14-14: Interpolation Search Parameters 75
   
Table 14-15: Interpolation Composite Restrictions 75
   
Table 14-16: Interpolation Composite Restrictions 76
   
Table 14-17: Comparison of OK grades with NN Grades within Resource Pit 76
   
Table 14-18: Base Case Recoveries, Metal Prices and Resulting Net Smelter Prices 92
   
Table 14-19: Measured Resource by Cu Equivalent Grade 93
   
Table 14-20: Indicated Resource by Cu Equivalent Grade 93
   
Table 14-21: Measured + Indicated Resource by Cu Equivalent Grade 94
   
Table 14-22: Inferred Resource by Cu Equivalent Grade 94
   
Table 15-1: Pit Slope Angle Recommendations 98
   
Table 15-2: Metallurgical Recoveries Used in Lerchs-Grossman Evaluations 100
   
Table 15-3: Base-case Lerchs-Grossman Economic Parameters 101
   
Table 15-4: Overall Slope Angles Used in Lerchs-Grossman Analyses 103
   
Table 15-5: Lerchs-Grossman Results – Metal Price Sensitivities 107
   
Table 15-6: Lerchs-Grossman Results – Cost Sensitivities 108
   
Table 15-7: Pit Design Parameters 111
   
Table 15-8: Ore Definition Parameters 119
   
Table 15-9: Proven Mineral Reserves by Phase 122
   
Table 15-10: Probable Mineral Reserves by Phase 122
   
Table 15-11: Combined Proven and Probable Mineral Reserves by Phase 122
   
Table 15-12: Combined Proven and Probable Mineral Reserves by Rock Formation 123
   
Table 16-1: Mine Production Schedule Criteria 124
   
Table 16-2: Mill Ramp-Up Schedule (Year 1) 124
   
Table 16-3: Annual Mill Throughput 125
   
Table 16-4: Mine Production Estimated Stripping Capacity Ramp-Up Schedule 127
   
Table 16-5: Waste Rock Storage Design Criteria 128
   
Table 16-6: Mine Production Schedule – Combined Proven & Probable Mineral Reserves 164
   
Table 16-7: Waste Rock Storage Tonnages by Destination 165
   
Table 16-8: Major Fleet Requirements 167
   
Table 16-9: Equipment Mechanical Availability 169
   
Table 16-10: Material Characteristics 170

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Table 16-11: Haul Cycle Times and Truck Productivities 172
   
Table 16-12: Equipment Operating Hours 174
   
Table 16-13: Hourly and Salary Employee Count 176
   
Table 18-1: Conceptual Summary of Groundwater Well Production 183
   
Table 18-2: Trip Data 192
   
Table 20-1: List of Agencies and Permits 217
   
Table 21-1: Capital Cost Summary by Area 227
   
Table 21-2: Sustaining Capital 229
   
Table 21-3: Life of Mine Operating cost Summary 230
   
Table 22-1: Total Mine Production Statistics 233
   
Table 22-2: Life of Mine Metal Production 234
   
Table 22-3: Smelter Return Factors 234
   
Table 22-4: Base Case and Historical Metals Prices 236
   
Table 22-5: Long Term Metals Prices 236
   
Table 22-6: Commodity Price Analysis 237
   
Table 22-7: Cash Copper Unit Cost Net of By Product Credits 238
   
Table 22-8: Net Income After Tax 240
   
Table 22-9: After Tax Economic Analysis – Combined Base Case (60/40) ($ millions) 240
   
Table 22-10: After Tax Economic Analysis – Combined Case – Historical 36 Month Prices 241
   
Table 22-11: After Tax Economic Analysis- Combined Case- Long Term Prices* 241
   
Table 22-12: Combined Base Case (60/40 split) 242

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LIST OF APPENDICES

APPENDIX DESCRIPTION
   
A Feasibility Study Contributors and Professional Qualifications
   
  •      Certificate of Qualified Person (“QP”) and Consent of Author
   
B Unpatented Claims List
   
C Patented Claims & Fee Land List
   
D LG Set 1 Cases
   
E Reserves Report
   
F Benches Mined by Period

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1                        SUMMARY

1.1                     PROPERTY

The Rosemont Copper Project is a planned copper mining development, containing appreciable molybdenum and silver by-products that is being developed by Augusta Resource Corporation (Augusta). The development is situated within the Rosemont Mining District on the northeastern flank of the Santa Rita mountain range and extends into the Helvetia Mining District on the western flank of the range. The Property consists of patented lode claims, unpatented lode claims, and fee lands comprising approximately 19,800 acres (8,012 hectares). Rosemont has also acquired 21 parcels that are more distal from the project area for infrastructure purposes, comprising an additional 300 acres (121 hectares).

Mining activity in the Helvetia and Rosemont Mining Districts dates to the mid-1800s, and by the 1880s production from mines on both sides of the Santa Rita Mountains supported the construction and operation of the Columbia Smelter at Helvetia, on the western side, and the Rosemont Smelter in the Rosemont Mining District on the eastern side. Production ceased in 1951 after production of about 227,300 tons of ore containing an estimated 17.3 million pounds of copper, 1.1 million pounds of zinc and 180,760 ounces of silver.

The copper mineralization of the Rosemont Deposit is primarily sulfide with an overlaying cap of oxide mineralization. The sulfide ore will be mined through conventional open pit mining techniques. Sulfide ore will be processed by crushing, grinding, and flotation to produce a copper concentrate product and a molybdenum concentrate product for market. This property will employ a conventional SAG mill/flotation circuit processing 75,000 short tons per day, equivalently indicated as either stpd or tpd.

1.2                     LOCATION

The Rosemont copper-molybdenum-silver deposit is located in Pima County, Arizona, USA on the northeastern flank of the Santa Rita Mountains approximately 30 miles southeast of the city of Tucson, Arizona. The Property occupies flat to mountainous topography at a surface elevation ranging from 4,000 feet to 6,290 feet and at geographical coordinates of approximately 31° 50’ N and 110° 45’ W.

1.3                     OWNERSHIP

Augusta signed an option agreement on the Rosemont Property in 2005. During the option period, Augusta completed a two-phase drilling program in 2005 and 2006. Augusta completed the purchase of a 100% interest in the property in March 2006. The purchase is subject to a 3% Net Smelter Return (NSR).

Augusta maintains offices in Denver, Colorado, USA, and Vancouver, British Columbia, Canada. The company’s common share are traded on the New York Stock Exchange MKT and the Toronto Stock Exchange under the symbol AZC.

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1.4                     GEOLOGY AND MINERALIZATION

The Rosemont Deposit is typical of the porphyry/skarn copper class of deposits. Similar to many other southwestern USA deposits in this class, Rosemont consists of broad-scale skarn mineralization developed in Paleozoic-aged carbonate sedimentary rocks adjacent to their contact with quartz-latite or quartz-monzonite porphyry intrusive rocks. Broadly disseminated sulfide mineralization occurs primarily in the altered Paleozoic skarn units and to a lesser extent in the altered intrusive units. Near surface weathering has resulted in the oxidation of the sulfides in the overlying Mesozoic units.

1.5                     EXPLORATION AND SAMPLING

Exploration of the Rosemont Deposit by previous companies consisted of 179 drill holes for a total of 210,200 feet. Since 2005, Augusta has drilled an additional 87 holes for a total of 132,500 feet. In 2005, Augusta carried out a 15-hole, 27,402-foot diamond drilling program. In 2006, Augusta completed a 40-hole, 68,727-foot diamond drilling program, consisting of resource, geotechnical, and metallurgical holes. Also in 2006, Augusta did extensive resampling and assaying of historic drill holes to fill-in missing data needed for resource calculations. In 2008, Augusta completed a 20-hole, 17,522-foot diamond drilling program, along with the sampling of ten geotechnical holes that had been drilled in 2006, but had not been sampled. Augusta recently completed a 12-hole, 18,649-foot diamond drilling program, along with the additional sampling of core from five older holes. The recent drilling included six holes (7,698 feet) drilled to collect metallurgical test samples, three exploration holes (5,466 feet) drilled to test geophysical targets, and three infill holes (4,711 feet) drilled in support of a revised resource calculation. The results of all of these drilling programs have been used to estimate the mineral resources presented in this report.

The older drilling was conducted by major companies using industry standard procedures of the time and has since been validated by Augusta under the direction of various Qualified Persons. The newer Augusta work has been conducted using standard industry protocols, including Quality Assurance/Quality Control procedures, all under the supervision of Qualified Persons. It is believed that the resulting drill hole database is reliable and can be confidently used in the estimation of the Rosemont resource and reserves.

Additional exploration conducted in 2011 included deep-penetrating induced polarization geophysical surveys (Titan 24). The results identified a number of anomalous responses that may be indicative of potential mineralization. During late 2011/early 2012 the western end of one of the anomalies was partially drill tested, intercepting variable mineralization near the top of the anomaly.

1.6                     MINERAL RESOURCE

The mineral resource estimation work was performed by Susan Bird, M.Sc., P. Eng. a Senior Associate at MMTS and an independent Qualified Person under the standards set forth by NI 43-101 (CIM, 2005). The resource is estimated using a 3-dimensional geologic model of all known lithologies and zones to create a block model encompassing the project area. The mineral resource estimates are effective as of July 17, 2012.

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Drill hole data including Cu, Mo and Ag grades is incorporated into the modeling by creating 50’ bench composites, corresponding to the planned bench height and elevations. Statistical and geostatistical analyses have been used to:

  1.

determine domain boundaries

     
  2.

determine the capping values used to restrict the outlier range of influence during interpolation,

     
  3.

determine the rotational and kriging parameters required for interpolation

     
  4.

determine appropriate sets of composites to use during interpolation that will preserve the tonnage-grade distribution of the data while allowing internal smoothing to account for dilution

In addition, several validation procedures have been performed on the Rosemont resource model. These checks include a comparison of mean grades as a global grade bias check, a set of swath plots to compare the nearest neighbor (NN) grades to the modeled Cu, Mo, and Ag grades, visual comparisons of drill hole assay and composite data with the modeled grades in section and plan, and verification of the change of support adjustment. Based on the results of this validation, it is the author’s opinion that the Rosemont resource model is globally unbiased and is appropriate for use in pit optimization and long range mine planning.

A Lerchs-Grossman (LG) pit shell having a 45-degree slope angle has been applied to the three dimensional block model to ensure reasonable prospects of economic extraction for the reported mineral resources. Metal prices used for the resource pit are $3.50/lb Cu, $15/lb Mo and $20/oz Ag. The mining costs used in the resource pit optimization for ore are $0.777/ton and for waste is $0.882/ton, with processing plus general and administration (G&A) costs of $4.90/ton for sulfide/mixed material and processing costs of $3.03/ton for oxide material. These costs are in line with those developed for use in the mineral reserves.

For the reporting of the in-situ resource by equivalent copper (EqvCu) within the LG pit shell, the metallurgic recoveries, metal prices, and resulting net smelter prices (NSPs) used, are summarized in Table 1-1.

Table 1-1: Base Case Recoveries, Metal Prices and Resulting Net Smelter Prices

Metal Metal
Price
Oxides Mixed Sulfide
NSP Recovery NSP Recovery NSP Recovery
Cu $2.50 /lb $2.425 /lb 65% $2.078 /lb 40% $2.078 /lb 86%
Mo $15 /lb 0 0 $13.095 /lb 30% $13.095 / lb 63%
Ag $20 /oz 0 0 $17.111 /oz 38% $17.111/oz 80%

The equivalent copper grades are calculated based on the above information, resulting in the following equations for each metallurgical zone:

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Sulfide: EqvCu% = Cu% + (Mo% * 0.63 * 13.095) + (AgOPT * 0.80 * 17.111)  
          (0.86 * 2.078)   (0.86 * 2.078 * 20)  
                 
Mixed: EqvCu% = Cu% + (Mo% * 0.30 * 13.095) + (AgOPT * 0.38 * 17.111)  
          (0.40 * 2.078)   (0.40 * 2.078 * 20)  
                 
Oxide: EqvCu% = Cu%          

The in-situ resource is classified as Measured, Indicated or Inferred corresponding to Canadian National Instrument 43-101 standards (CIM, 2005). The resource by equivalent copper grade for the Rosemont Deposit is summarized in Table 1-2 for Measure, Indicated, Measured+Indicated, and Inferred mineral resources, along with the base case equivalent copper values for each zone (oxide, mixed, sulfide). These cutoffs are sufficient to cover the processing plus G&A costs for the sulfide and mixed material ($4.90/ton) and the processing costs of the oxide material ($3.03/ton), at the expected metallurgical recoveries.

The measured and indicated mineral resource presented here is inclusive of the mineral reserve presented in the Mineral Reserve section. Mineral resources that are not mineral reserves do not have demonstrated economic viability.

Due to the uncertainty that may be associated with Inferred mineral resources it cannot be assumed that all or any part of inferred mineral resources will be upgraded to an Indicated or Measured resource.

Table 1-2: Base Case Mineral Resource by Classification and Zone

Class Zone Tons
(millions)
Cu Eqv
(%)
Cu (%) Mo (%) Ag (opt)
Measured Oxide
Mixed
Sulfide
30.3
13.1
334.6
0.17
0.64
0.56
0.17
0.59
0.44
---
0.006
0.015
---
0.05
0.12
Indicated Oxide
Mixed
Sulfide
33.1
36.8
534.7
0.16
0.56
0.48
0.16
0.51
0.37
---
0.007
0.014
---
0.05
0.11
Measured+Indicated Oxide
Mixed
Sulfide
63.4
50.0
869.4
0.17
0.58
0.51
0.17
0.53
0.40
---
0.007
0.014
---
0.05
0.11
Inferred Oxide
Mixed
Sulfide
1.1
10.1
128.5
0.15
0.43
0.49
0.15
0.39
0.40
---
0.006
0.013
---
0.02
0.10

Base Case cutoff grades: oxide 0.10%CuEqv, mixed 0.30% CuEqv, sulfide 0.15%CuEqv.

Augusta’s 2012 drilling campaign at the Rosemont Deposit has increased both the quantity and confidence level of the estimated mineral resources, which presently totals about 919.3 million tons of measured and indicated, sulfide and mixed mineral resources grading 0.51% CuEqv, 0.41% Cu, 0.014% Mo, and 0.11 ounces per ton Ag, at a 0.15% CuEqv cutoff for sulfide and 0.30% CuEqv cutoff for a minor mixed component. An additional 138.6 million tons of inferred sulfide and mixed mineral resources are estimated at a grade of 0.49% CuEqv, 0.40% Cu, 0.012% Mo, and 0.10 ounces per ton Ag, at the same cutoffs. Sulfide and mixed material can be combined as metallurgical testwork of the mixed material indicates that it can be processed with the sulfide material to produce a concentrate. Augusta’s recent drilling program and resource modeling was successful in converting significant tonnages of material previously classified as inferred into measured and indicated resource.

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In addition, geologic and metallurgical studies conducted by Augusta have shown the potential for considering the oxide copper mineralization that overlies the sulfide deposit. Estimated measured and indicated oxide mineral resources total 63.4 million tons grading 0.17% Cu, at a 0.10% CuEqv cutoff (for oxide % CuEqv = % Cu). An additional inferred oxide mineral resource of 1.1 million tons grading 0.15% Cu is present, using the same cutoff. Oxide material could potentially be processed by heap leaching to recover the copper.

The classification of currently inferred sulfide and oxide mineral resources can potentially be improved with further drilling. Additional mineral resources may be found in extensions to the north and down-dip of the Rosemont Deposit. Mineralization is also known to occur at Broadtop Butte, which could potentially be added as a satellite development. Further mineralization also occurs in the Copper World and Peach-Elgin deposits on the Rosemont Property.

1.7                     MINE RESERVES & MINE PLAN

The Rosemont Deposit is a large tonnage, copper-molybdenum deposit located in close proximity to the surface and amenable to open pit mining methods. The proposed pit operations will be conducted from 50-foot high benches using large-scale mining equipment.

The mine has a 21-year life, with sulfide ore to be delivered to the processing plant at an initial rate of 75,000 tpd. Provisions are included to increase production to 90,000 tons of ore per day (tpd) in year 12 of operations. Seven mining phases have been defined for the extraction sequence for the Rosemont Deposit. The phase development strategy consists of extracting the highest metal grades along with the minimum strip ratios during the initial years to maximize the economic benefits of the ore-body.

The mineral reserve estimates presented in this report were prepared by Mr. Robert Fong, P.Eng., Principal Mining Engineer for Moose Mountain Technical Services. Mr. Fong meets the requirements of an independent Qualified Person under NI 43-101 standards. The mineral reserve estimates are effective as of July 24, 2012.

Proven and probable mineral reserve estimates and waste rock for the Rosemont Deposit are summarized by mining phase in Table 1-3.

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Table 1-3: Combined Proven and Probable Mineral Reserves by Phase

Phase Sulfides >= 4.90 $/ton NSR Cut-off
Waste
Ktons
Total
Material
Ktons

Strip
Ratio

Ktons

NSR $/t

TCu %
Mo %
Ag oz/t
                 
1 61,546 22.38 0.50 0.016 0.14 142,729 204,275 2.32
2 27,169 17.24 0.40 0.011 0.09 84,526 111,695 3.11
3 42,418 19.37 0.40 0.020 0.13 59,553 101,971 1.40
4 42,699 21.54 0.49 0.013 0.14 100,709 143,408 2.36
5 79,845 21.64 0.50 0.013 0.13 156,603 236,448 1.96
6 241,477 17.99 0.42 0.014 0.12 411,973 653,450 1.71
7 172,052 19.16 0.42 0.015 0.11 287,362 459,414 1.67
                 
Total 667,206 19.42 0.44 0.015 0.12 1,243,455 1,910,661 1.86
                 

(NSR values are based on metal prices of $2.50/lb Cu, $15.00/lb Mo and $20.00/oz Ag.)

The pit design reflects an optimum pit at metal price of $1.88 /lb Cu, $11.07 /lb Mo, and $14.87/oz Ag. Proven and probable sulfide mineral reserves within the designed final pit total 667 million tons grading 0.44% Cu, 0.015% Mo and 0.12 oz Ag/ton. There are 1.24 billion tons of waste materials, resulting in a stripping ratio of 1.9:1 (tons waste per ton of ore). Total material in the pit is 1.9 billion tons. Contained metal in the sulfide (proven and probable) mineral reserves is estimated at 5.88 billion pounds of copper, 194 million pounds of molybdenum and 80 million ounces of silver. No mineralized oxide materials are in the ore reserves, they are included with the waste materials.

Nearly 46% of the sulfide mineral reserves in the Rosemont ultimate pit are classified as proven and the remainder (54%) is considered probable. The classifications are based on the exploration drilling in the Rosemont deposit. All of the mineral reserve estimates reported above are contained in the mineral resource estimates presented in Section 14.

The Rosemont ultimate pit contains approximately 24 million tons of inferred sulfide mineral resources that are above the $4.90/ton NSR cutoff value for sulfides. These resources are included in the waste estimates presented in Table 1-3. Inferred mineral resources are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves. Inferred mineral resources have a great amount of uncertainty as to their existence and as to whether they can be mined economically. It cannot be assumed that all or any part of inferred mineral resources will ever be upgraded.

1.8                     MINERAL PROCESSING & METALLURGICAL TESTING

The earliest existing records of metallurgical testing are from the period 1974 - 1975, at which time grinding and flotation tests were performed. In the first half of 2006, Augusta initiated test work to provide a better understanding of the metallurgy of the Rosemont Deposit and establish the design criteria for the design of a process facility. Additional test work was conducted in 2012 on new drill core obtained from the Augusta drilling program in December 2011.

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The representative ore samples were tested to determine grinding and flotation criteria. The test work indicates a process of crushing and grinding the ore to 80% passing 105 micron size distribution followed by bulk flotation to recover copper and molybdenite minerals. A molybdenite concentration circuit to treat the bulk flotation concentrate will be able to produce a molybdenite concentrate.

An estimate of metal production in concentrate for the first 21 years of plant operation was prepared from the results of flotation test work. The estimates of annual metal recovery are presented in Table 1-4 .

Table 1-4: Estimated Metal Recovery by Year of Production

Estimated Metal Recovery by Year of Production

Production Year
Recovery %
Cu Mo Ag
       
1 89.8 65.0 77.5
2 89.8 65.0 77.5
3 89.8 65.0 77.5
4 84.1 34.2 72.6
5 84.1 34.2 72.6
6 84.1 34.2 72.6
7 84.1 34.2 72.6
8 90.6 78.7 78.2
9 90.6 78.7 78.2
10 84.8 74.3 73.9
11 82.1 72.2 71.8
12 84.4 73.9 73.5
13 84.0 56.7 73.1
14 85.5 57.2 74.3
15 89.1 58.6 76.9
16 89.1 58.6 76.9
17 89.1 58.6 76.9
18 89.1 58.6 76.9
19 89.1 58.6 76.9
20 89.1 58.6 76.9
21 89.1 58.6 76.9

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1.9                     RECOVERY METHODS

Sulfide ore will be transported from the mine to the primary crusher by off-highway haulage trucks then conveyed to the concentrator facilities. Copper concentrate produced at the concentrator facility will be loaded into highway haul trucks and transported to a concentrate smelter and metal refinery. Molybdenum concentrate produced at the concentrator facility will be bagged and loaded onto trucks for shipment to market.

The process selected for recovering the copper and molybdenite minerals can be classified as “conventional”. The sulfide ore will be crushed and ground to a fine size and processed through mineral flotation circuits.

1.10                    ENVIRONMENTAL

Permitting for the Rosemont Copper Project involves federal approvals and requires compliance with the National Environmental Policy Act (NEPA). This in turn requires an Environmental Impact Statement (EIS) and compliance with the Endangered Species Act (ESA) and the National Historic Preservation Act (NHPA). A Mine Plan of Operation was submitted to the US Forest Service on July 11, 2007 to initiate the EIS and start the permitting process. Major federal permits required to construct and begin operation of the Rosemont Project includes an Environmental Impact Statement Record of Decision from the USFS permitting the use of Federal lands and a Clean Water Act (CWA) Section 404 permit for discharge of fill material to on-site washes. Major state permits include an Aquifer Protection Permit, a 401 Certification, and an Arizona Pollution Discharge Elimination System (AZPDES) general storm water permit. The only major local permit required is a Pima County Clean Air Act (CAA) Title V air quality permit. Other permits which do not affect the timeline for project permitting and subsequent start up include explosives permits, nuclear instrumentation licenses, hazardous waste identification, tracking numbers and spill control plans. A list of permits is provided in Section 20.

1.11                    ECONOMIC ANALYSIS

1.11.1                  Operating Costs

The mine operating costs were derived from equipment hours and cycle times developed by Moose Mountain from their mine plan. Rebuild costs for major equipment were generated from vendor supplied component replacement schedules and URS Energy and Constructions’ data base for similar projects and equipment. Mining costs supplied by others were checked by URS who built the estimate and is the responsible Qualified Person. The average life of mine operating costs for the mining operation is $1.23 per ton mined. These costs include: clearing of vegetation, removal of topsoil, drilling, blasting, loading, hauling, road and dump maintenance, re-grading, mine operations supervision, craft labor and subcontractor costs.

Mill process operating costs for the life of mine average $4.27/ton of mill ore which includes crushing and conveying, grinding and classification, flotation and regrind, concentrate thickening, filtration and dewatering, tailings disposal and mill ancillary services.

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The average life of mine operating cost for the supporting facilities and general administrative expenses is $0.59/ton of mill ore. The supporting facilities include laboratory, safety and environmental, accounting, human resources, security and the general manager’s office. Also included is an endowment, railcar lease, and CAP water.

The life of mine average site direct operating cost estimate by cost center is shown in Table 1-5 below. All costs are estimated in second quarter 2012 Dollars at an accuracy of ± 10%.

Table 1-5: Summary of Average Life of Mine Operating Costs

  Annual Cost ($000)
Mining $106,000
Mill Operations & Maintenance $134,407
Support Facilities and G&A $18,484
   
Total $258,891

1.11.2                Initial Capital Cost

The total capital cost estimate to design, construct and commission the Rosemont facilities is estimated to be $1,060.4 million for the sulfide plant. The estimate includes the direct field cost for constructing the project at $870.6 million as well as $189.8 million for the indirect costs associated with the design engineering, procurement and construction, commissioning, spare parts, contingency, power line gross-up tax, and excludes Owner’s cost. All costs are expressed in second quarter 2012 Dollars at an accuracy of ± 10% with no allowance provided for escalation, interest, foreign currency, hedging, or financing during construction.

1.11.3                Financial Analysis

The Rosemont Project economics were prepared using a discounted cash flow model. Costs are in constant second quarter 2012 Dollars with no provisions for escalation. The financial indicators examined for the project included the Net Present Value (NPV), Internal Rate of Return (IRR) and payback period (time in years to recapture the initial capital investment). Annual cash flow projections were estimated over the life of the mine based on capital expenditures, production costs, transportation and treatment charges and sales revenue. The life of the mine is 21 years.

The sales revenue is based on the production of three commodities: copper, molybdenum and silver. Gold is also present in the copper concentrates in the form of a saleable by-product credit. The estimates of capital expenditures and site production costs have been developed specifically for this project.

Metal sales prices used in the evaluation are listed in Table 1-6.

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Table 1-6: Base Case and Historical Metals Prices


60/40 Weighted
Average *
3 Year Historical
Average
Copper $3.50 / pound $3.56 / pound
Molybdenum $14.19 / pound $15.06 / pound
Silver $3.90 / ounce $3.90 / ounce
Gold $450.00 / ounce $450.00 / ounce

*60/40 weighted average of the 36 month historic price and the 24 month futures price forecast Silver and gold metal prices are set from the Silver Wheaton agreement

In addition to the above metal sales price cases, a case of long term metal prices was also evaluated. Long term metal prices are shown in Table 1-7 below.

Table 1-7: Long Term Metals Prices

Copper $2.62/lb
Molybdenum $15.00/lb
Silver $3.90/oz
Gold $450.00/oz

The after-tax financial results for the three metal pricing scenarios are shown in Table 1-8.

Table 1-8: Financial Indicators (After Tax)



Base Case
(60/40
split)
Historical
36
Months
Long Term
Metal
Prices
NPV 0% $7,257.5 $7,498.4 $4,554.4
NPV 5% $3,645.8 $3,776.4 $2,256.0
NPV 8% $2,507.6 $2,603.1 $1,529.4
IRR 37.9% 38.8% 30.9%
Payback Years 2.3 2.2 2.4

1.12                    CONCLUSIONS & RECOMMENDATIONS

The Rosemont Deposit resource and mineral reserves have increased with the additional drilling campaign in 2012. Metallurgical recoveries improved slightly for copper with the additional metallurgical testing. Metal prices have improved since the 2009 feasibility study update; however, silver and gold prices used in this update are lower reflecting an agreement with Silver Wheaton for a forward sale of gold and silver. The after-tax NPV, IRR, and payback indicators are also improved over the previous 2009 update.

With the improved economic indicators, the Rosemont Copper Project should continue with the design engineering and construction of the facilities as soon as the permitting effort allows.

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2                        INTRODUCTION

2.1                     GENERAL

Augusta Resource Corporation (Augusta) is a base metals company with its corporate office located in Vancouver, British Columbia, Canada, and executive office located in Denver, Colorado. Augusta is focused on advancing its Rosemont Copper Project near Tucson, Arizona. The project is nearing the last stages of permitting with the Record of Decision (ROD) expected to be received in the fourth quarter of 2012 and production to start-up in 2015. Augusta’s objective is to build and operate the world-class Rosemont mine and develop a robust portfolio of assets in North America with the focus on organic growth and early stage acquisitions. Augusta trades on the Toronto Stock Exchange and the NYSE MKT under the symbol AZC. Rosemont Copper Company (Rosemont) is a wholly owned subsidiary of Augusta Resource Corporation and will be the operating company for the mine and process facilities.

In 2006, Augusta Resources Corporation retained a number of contractors, including M3 Engineering & Technology Corporation (M3), to provide a review of prior work on the Rosemont Copper Project and prepare technical and cost information to support a bankable level Feasibility Study and Technical Report compliant with the Canadian National Instrument (NI) 43-101 and Form 43-101F1. The Technical Report titled “NI 43-101 Technical Report for the Rosemont Copper Project Feasibility Study, Pima County, Arizona, USA” was issued in August 2007. An update of the 2007 Technical Report, titled “NI 43-101 Technical Report for the Rosemont Copper Project Updated Feasibility Study, Pima County, Arizona, USA”, incorporating additional resource information and metallurgical testing, was issued in January 2009 and amended in March 2009.

2.2                     PURPOSE OF REPORT

The purpose of this technical report is to present updated mineral resource information and metallurgical testing information completed since the 2009 technical report update. Capital costs, operating costs, and the economic analysis were updated to 2012 costs based on 40% completion of engineering. Basic engineering was completed in November 2010. It is Augusta’s intent to continue to develop the Rosemont Copper Project once the Record of Decision (ROD) has been received from the US Forest Service.

2.3                     SOURCES OF INFORMATION

This report is based on data supplied by Augusta and Rosemont and information developed during the feasibility study and basic engineering period by M3 and other third party consultants. The source documents are summarized is Section 27.

2.4                     CONSULTANTS AND QUALIFIED PERSONS

Augusta retained a number of contractors, including M3 Engineering and Technology Corporation (M3), to provide a review of prior work on the project and prepare technical and cost information to support an updated Feasibility Study and this Technical Report following the outline as defined in Canada National Instrument (NI) 43-101 and in compliance with Form 43-101F1. Dr. Conrad Huss, P.E. of M3 Engineering and Technology Corporation (M3) is the Principal Author and Qualified Person responsible for the preparation of this report. Dr. Huss has visited the site on numerous occasions prior to this Updated Feasibility Study and is familiar with the site. In addition, the following M3 employees, under the supervision of Dr. Huss, visited the site during the initial Feasibility Study and/or the 2009 Updated Feasibility Study on the dates noted.

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  • Thomas L. Drielick, P.E., Senior Vice President; August 21, 2007
  • Rex Henderson, P.E., Project Manager; August 9, 2006, September 27, 2006 and June 12, 2007
  • David Moll, P.M.P., Asst. Project Manager, November 13, 2008
  • Randy Hensley, Construction Manager, June 12, 2007
  • Enrico Laos, P.E., Electrical Supervisor; August 9, 2006
  • Tim Oliver, P.E., Environmental Specialist; September 27, 2006
  • Daniel Roth, P.E., Reclamation Lead; May 25, 2007, July 11, 2007, November 13, 2008
  • Craig Hunt, Civil Lead, November 13, 2008
  • David Caldwell, Civil Designer, May 25, 2007 and July 11, 2007
  • Tony Ottinger, Civil Designer, May 25, 2007 and July 11, 2007
  • Robert Davidson, Project Engineer, August 9, 2006
  • Francisco Espinosa, Civil Designer, August 9, 2006

Other contributing authors and Qualified Persons responsible for preparing this Updated Feasibility Study Technical Report include; Ms. Susan C. Bird, M.Sc., P. Eng., of Moose Mountain Technical Services; Mr. Robert H. Fong, P. Eng., of Moose Mountain Technical Services; Mr. John I. Ajie, P.E., of URS Energy and Construction; and Mr. Thomas L. Drielick, P.E., of M3 Engineering and Technology Corporation.

Augusta retained Moose Mountain Technical Services (MMTS) to develop and oversee the resource estimate, mineral reserve estimate and mining methods. Ms. Susan C. Bird, M.Sc., P.Eng., Senior Associate with MMTS, authored Section 14 – Mineral Resource Estimate and Section 23 – Adjacent Properties. Ms. Bird also reviewed Section 7 – Geological Setting and Mineralization, Section 8 – Deposit Types, Section 9 – Exploration, Section 10 – Drilling, Section 11 – Sample Preparation, Analysis and Security, and Section 12 – Data Verification. These sections were authored by Mr. Mark Stevens of Augusta Resource Corporation and taken from the previous Updated Feasibility Study. Ms. Bird is the Qualified Person responsible for these sections of the current report. Ms. Bird visited the site from January 30, 2012 to February 3, 2012.

Mr. Robert H. Fong, P.Eng., Principal Mining Engineer with Moose Mountain Technical Services is the Qualified Person responsible to estimate and oversee the calculations of the open pit reserves and to develop the Life of Mine (LOM) Mine Plan which includes a Lerchs-Grossman analysis, pit design, mine production schedule, mine access and haul roads and waste rock stockpiles. Mr. Fong authored Section 15 – Mineral Reserve Estimate and Section 16 –Mining Methods (16.1 to 16.6) and is the Qualified Person responsible for these sections of the current report. Mr. Fong visited the project site on November 20, 2008.

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Mr. John I. Ajie, P.E. – Vice President of Engineering Civil Construction & Mining Group, with URS Energy and Construction is the Qualified Person responsible for supervising and reviewing the development of the mine capital and operating cost estimate. Mr. Ajie is responsible for the basis for the mine capital and operating costs in Sections 16.7 to 16.14 and mine capital and operating costs in Section 21. Mr. Ajie visited the site on May 18, 2006.

Mr. Thomas L. Drielick, P.E., Senior Vice President of M3 Engineering & Technology Corporation, is the Qualified Person responsible for reviewing the metallurgical test work and establish the process recoveries and process recovery methods. Mr. Drielick visited the site on August 21, 2007.

M3 Engineering and Technology Corporation (M3) of Tucson, Arizona was retained by Augusta to prepare the process and infrastructure design, capital and operating costs for the process and infrastructure, and integrating the work by other consultants into this Updated Feasibility Study including the overall project capital cost estimate, operating cost estimate, implementation schedule for the project, and an economic analysis. M3 also reviewed previous metallurgical test reports and coordinated additional metallurgical testing programs conducted by SGS Lakefield Research Limited (SGS) of Toronto, Ontario, and Vancouver, British Columbia, Canada; Mountain States Research & Development Inc. (MSRDI) of Tucson, Arizona; Hazen Research, Inc. (HRI) of Golden, Colorado; and G&T Metallurgical Services (G&T) of Kamloops, British Columbia, Canada, all under contract with Rosemont. SGS Lakefield was contracted to conduct ore grindability characterization tests and establish a preliminary grinding circuit design utilizing Comminution Economic Evaluation Tool (CEET) software. MSRDI, SGS, and G&T were contracted to conduct batch and locked cycle flotation tests to define ore variability, grind / grade / recovery parameters, and reagent screening to define a reagent scheme. MSRDI also conducted dewatering tests for concentrate and tailings. Hazen Research was contracted to conduct Bond rod and ball mill index tests. G&T Metallurgical Services Ltd. was contracted to assess mineral content, mineral liberation, and association and mineral fragmentation characteristics on two ore samples from MSRDI. The SGS Lakefield report, MSRDI report and Hazen report are referenced in this Technical Report and formed the basis for establishing the plant design parameters, concentrate grades, metal recoveries, mill sizing and reagent consumptions.

2.5                     DEFINITION OF TERMS USED IN THIS REPORT

The units of measure in this report are US units and all costs are in US Dollars, unless otherwise noted. The unit of mass is the short ton (ton, T, or t). A short ton is 2,000 pounds. Other units used include dry ton (DT, dt), miles (mi), feet (ft.), inches (in), acres (ac), square feet (ft2, sq. ft.), square inch (in2, sq. in.), cubic feet (ft3, cu. ft.), gallon (g), gallons per minute (gpm), pound (lb., lbs.), pound per ton (lb./t), Fahrenheit temperature (° F), year (Y, y), day (D, d), hour (h), minutes (m) and seconds (s). Silver and gold quantities and grade are in troy ounces (oz.) and troy ounces per ton (opt), respectively.

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Acronyms and abbreviations used in this report are noted below:

  AA Atomic Absorption Spectrometry
  AAC Arizona Administrative Code
  ACC Arizona Corporation Commission
  ADEQ Arizona Department of Environmental Quality
  Ag Silver
  Anaconda Anaconda Mining Company
  Anamax Anamax Mining Company
  ANPL Arizona Native Plant Law
  APP Aquifer Protection Permit
  ARS Arizona Revised Statutes
  ASARCO American Smelting and Refining Company
  Au Gold
  Augusta Augusta Resource Corporation
  AZPDES Arizona Pollutant Discharge Elimination System
  BADCT Best Available Demonstrated Control Technology
  Banner Banner Mining Company
  BLM Bureau of Land Management
  BMP Best Management Practices
  CAA Clean Air Act
  CAP Central Arizona Project
  CESQG Conditionally Exempt Small Quantity Generators
  CFR Code of Federal Regulations
  CGP Construction General Permit
  CLS Conservation Land System
  Cu Copper
  CuEqv Copper Equivalent
  CWA Clean Water Act
  EIS Environmental Impact Statement
  EPA Environmental Protection Agency
  ESA Endangered Species Act
  G&T G&T Metallurgical Services
  HAP Hazardous Air Pollutants
  HRI Hazen Research Incorporated
  IP Individual Permit
  IRA Important Riparian Area
  kWh Kilowatt Hour
  LOM Life of Mine
  LQHUW Large Quantity Handlers of Universal Wastes
  M3 M3 Engineering and Technology Corporation
  MMTS Moose Mountain Technical Services
  Mo Molybdenum
  MSGP Multi-Sector General Permit
  MSRDI Mountain States Research and Development, Inc.
  MW Megawatts

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  NAAQS National Ambient Air Quality Standards
  NAVD 88 North American Vertical Datum 1988
  NEPA National Environmental Policy Act
  NHPA National Historic Preservation Act
  NOI Notice of Intent
  NPDES National Pollutant Discharge Elimination System
  NSPS New Source Performance Standards
  NSR Net Smelter Return
  NWP Nation Wide Permit
  QA/QC Quality Assurance and Quality Control
  PAH Pincock, Allen & Holt, Inc.
  PCDEQ Pima County Department of Environmental Quality
  RCRA Resource Conservation and Recovery Act
  RQD Rock Quality Data
  SGS SGS Lakefield Research Limited or SGS Vancouver
  Skyline Skyline Assayers and Laboratories, Inc.
  SQG Small Quantity Generators
  SQHUW Small Quantity Handlers of Universal Wastes
  SRM Standard Reference Material
  Stantec Stantec Consulting, Inc.
  SWPPP Storm Water Pollution Prevention Plan
  SWTC South West Transmission Cooperative
  SX-EW Solvent Extraction - Electrowinning
  TCLP Toxic Characteristic Leaching Procedures
  TCP Traditional Cultural Properties
  TCu Total Copper Concentrations
  TEP Tucson Electric Power
  TPD Tons Per Day
  USFWS US Fish & Wildlife Service
  UTM NAD 83 Universal Transverse Mercator – North American Datum 1983
  WAPA Western Area Power Administration
  Wardrop Wardrop Consultants
  WECC Western Electricity Coordinating Council
  WGI Washington Group International
  Winters The Winters Company
  WLRC WLR Consulting, Inc.
  XRF X-Ray Fluorescence

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3                        RELIANCE ON OTHER EXPERTS

Mr. Ron Hamagami of URS Energy and Construction provided capital and operating cost information for the mine in Section 21 – Capital and Operating Costs. Mr. Ajie reviewed this work and is the Qualified Person responsible for this section.

Mr. Brian Lindenlaub of WestLand Resources, Incorporated authored Section 18 – Project Infrastructure and Section 20 – Environmental Studies, Permitting and Social or Community Impact. M3 also relied on the information provided by Mr. Robert Loewen who authored Section 19 – Market Studies and Contracts and Mr. Mark Stevens of Augusta Resource Corporation who authored Section 4 – Property Description and Location. These sections were reviewed by M3 and judged to be professionally sound and to industry standards.

Mr. David Nicholas of Call & Nicholas Incorporated (CNI – Tucson, Arizona) prepared a slope stability study for the pit walls and prepared run of mine (ROM) fragmentation analysis for sulfide and oxide ore and the waste rock. Mr. Robert Fong, Principal Mining Engineer of MMTS, has reviewed and incorporated the CNI work into the mine design sections of this report.

Mr. Mark Stevens, C.P.G., Vice President of Exploration for Augusta, compiled the drill hole data files and prepared the geology section of this Updated Feasibility Study based on earlier published reports and internal reports (Augusta – 2007-2009). Mr. Stevens has spent time on site on numerous occasions over the last several years. Ms. Susan C. Bird of Moose Mountain Technical Services has review this section and is the independent Qualified Person responsible for this work.

Tetra Tech, Inc. of Golden, Colorado and Tucson, Arizona were responsible for the site geotechnical investigations consisting of a site geotechnical study, a geologic hazards assessment, and a baseline geochemical characterization study. Tetra Tech also provided a site water management plan, waste management plan, initial dry stack tailings facility design, and oxide leach facilities. Tetra Tech provided the design and material quantities for the storm water pond and compliance point dam. M3 estimated the capital cost based on the material quantities. The Tetra Tech reports are referenced in the Technical Report in Section 27.

Tetra Tech was also responsible for preparation of the reclamation and closure plan with some support from M3. Tetra Tech developed the concurrent reclamation plan and soil salvage estimates for the operational and storage areas of the site. Tetra Tech and Augusta estimated the annual costs for reclamation. Tetra Tech was also responsible for the Aquifer Protection Permit, supported by Errol L. Montgomery of Tucson, Arizona, who prepared the ground water model to confirm the impact of the project on the ground water.

AMEC of Denver, Colorado and Tucson, Arizona provided the design and material quantities for the dry stack tailings facility and process water temporary storage pond and M3 estimated the capital cost based on the material quantities.

Errol L. Montgomery & Associates (ELM) of Tucson, Arizona was responsible for the ground water hydrology modeling and studies to support Tetra Tech with the Aquifer Protection Permit. ELM was also responsible for the exploration drilling and testing of water wells to locate a system of wells to supply fresh water for the project. ELM provided the production well cost and design up to the well head.

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Stantec Consulting, Inc. of Tucson, Arizona was responsible for the conceptual design of the fresh water pipeline from the well fields to the project site. CDM Smith Inc. of Phoenix, Arizona, was responsible for the final design of the fresh water system and pumping stations, including a water surge analysis for the system. CDM Smith provided the design, the quantity take-offs for construction, and the capital cost estimate for the system.

The primary Qualified Persons responsible for preparing this Technical Report relied on the various reports and documents listed in Section 27. These reports and documents were prepared by technically qualified and professional persons and were found to be generally well organized, to industry standards, and where applicable, the conclusions reached were judged to be professionally sound. It is assumed that the information and explanations given to the Qualified Persons and those assisting the Qualified Persons by the employees of Augusta and third party consultants, who provided the reports referenced in Section 27 during the preparation of this Rosemont Copper Project Updated Feasibility Study and this Technical Report, were essentially complete and correct to the best of each employee’s or consultant’s knowledge and that no information was intentionally withheld.

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4                        PROPERTY DESCRIPTION AND LOCATION

4.1                     LOCATION

The Rosemont Property is located approximately 30 miles (50 km) southeast of Tucson, in Pima County, Arizona (Figure 4-1). The Property consists of a comprehensive land package that covers much of the Rosemont and Helvetia Mining Districts, occurring on the eastern and western sides, respectively, of the Santa Rita Range. The lands are under a combination of private ownership by Augusta and Federal ownership. The lands occur within Townships 18 and 19 South, Ranges 15 and 16 East, Gila & Salt River Meridian. The Rosemont Property geographical coordinates are approximately 31º 50’N and 110º 45’W.

Figure 4-1: Property Location of Rosemont Project

4.2                     LAND TENURE

On March 31, 2006, Augusta completed the purchase of a 100% interest in the Property for a total of US$20.8 million and continues to maintain the property in good standing. Augusta retained the legal firm of Fennemore Craig P.C. to handle the legal transfer of the Rosemont Property. Augusta’s land information has come from 2006 property purchase legal documents and has been subject to further validation contracted by Augusta, including a mining claim specialist, Daniel Mead of Tucson, Arizona, and registered mining claim surveyors at Darling Environmental & Surveying, Ltd. of Tucson, Arizona. Darling Environmental & Surveying, Ltd. has conducted an extensive field and office review of the patented and unpatented claims. Fennemore Craig has continued to have legal involvement with land title maintenance.

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The Rosemont Property is a combination of fee land, patented mine and mill site claims, and unpatented mine and mill site claims. Taken together, the land position is sufficient to allow mining of the open pit, processing of ore, storage of tailings, disposal of waste rock, and operation of milling equipment. These lands are accessible under the provisions of the Mining Law of 1872, subject to approval from the US Forest Service after the completion of an Environmental Impact Statement (EIS) as per the National Environmental Policy Act (NEPA) process. The EIS process includes interagency consultation on project alternatives and mitigation of environmental impacts. The use of the project surface rights requires obtaining a number of federal, state, and local permits and approvals, some of which are complete, others are in progress.

The core of the Rosemont Property consists of 132 patented lode claims that in total encompass an area of 2,000 acres (809 hectares) as shown in Figure 4-2. Surrounding the patented claims are a contiguous package of 1,060 unpatented lode-mining claims with an aggregate area of more than 16,000 acres (6,475 hectares). Most of the unpatented claims were staked on Federal land administered by the US Forest Service, but a limited number of claims in the northeast portion of the property are on Federal land administered by the Bureau of Land Management. Associated with the property are 33 parcels of fee (private) land consisting of 1,800 acres (728 hectares). The area covered by the patented claims, unpatented claims and fee lands totals approximately 19,800 acres (8,012 hectares). Rosemont has also acquired 15 parcels that are more distal from the project area that are planned for various infrastructure purposes including, well fields, pump stations, utilities, and ranch operation, comprising an additional 300 acres (121 hectares). A listing of the unpatented claims, patented and fee lands is provided in Appendix B and C, respectively.

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Figure 4-2: Rosemont Property Ownership

The patented lode claims and fee land parcels are both considered to be private lands that provide the owner with both surface and mineral rights. The patented mining claim block, located in the core of the property, is monumented in the field by surveyed brass caps on short pipes cemented into the ground. The fee lands are located by legal description recorded at the Pima County Recorder’s Office. The patented claims and fee lands are subject to annual property taxes amounting to a total of approximately $55,000. As long as the property taxes are paid annually on these claims, there is no expiration date.

US Forest Service and Bureau of Land Management lands have had the mineral rights reserved to Augusta on the unpatented lode mining claims that surround the patented claims. Wooden posts and stone cairns mark the unpatented claim corners, end lines and discovery monuments, all of which have been surveyed. The unpatented lode claims have no expiration and are maintained through the payment of annual maintenance fees of $140.00 per claim, for a total of about $150,000, payable to the Bureau of Land Management.

A 3% Net Smelter Return (NSR) royalty applies to the patented claims, most of the unpatented claims, and some of the fee land.

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5

ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

5.1                     ACCESSIBILITY

The Rosemont Copper Project is located in Pima County, Arizona, approximately 30 miles southeast of Tucson. The project site is accessible from Tucson by going east on Interstate Highway I-10 to Arizona State Route 83 and south on SR-83 approximately 11 miles to the main access road to the plant. The intersection of the proposed primary access road to the plant and SR-83 will be between mile markers 46 and 47 at a point that provides a clear line of site for up to 2,500 feet in each direction. SR-83 will be modified to provide safe ingress and egress from the main access road. Modifications will include a 500 foot long center lane in each direction for accelerating and decelerating. A 220 foot deceleration lane and 500 foot acceleration lane will also be constructed on the southbound lane of SR-83 for safe access into and out of the plant site. The primary access road into the plant is approximately 3.2 miles long and will be paved.

A secondary gravel access road to the plant will be provided from South Santa Rita Road on the west side of the Santa Rita Mountains over the mountain ridge to the plant. Access will be from Interstate Highway I-19 at Sahuarita then east on Santa Rita Road to the start of the west plant access road. The main power transmission line to the plant and the fresh water pipeline will generally follow the alignment along Santa Rita Road and the west access road. The west access road is approximately 4.4 miles long from Helvetia road into the plant and will be used to access the well fields and fresh water booster stations located in the Santa Cruz Valley to the west.

The city of Tucson, Arizona, provides the nearest major railroad and air transport services to support the project.

5.2                     CLIMATE

The southern Arizona climate is typical of a semi-arid continental desert with hot summers and temperate winters. The project area is at the north end of the Santa Rita Mountain Range at elevations between 4,550 feet and 5,300 feet above mean sea level (amsl). The higher elevation in the project area results in a milder climate than at the lower elevations across the region. Summer daily high temperatures are above 90 degrees Fahrenheit (°F) with significant cooling at night. Winter in the project area is typically drier with mild daytime temperatures and overnight temperatures that are typically above freezing. Winter can have occasional low intensity rainstorm patterns that can last for multiple days.

The average annual precipitation in the project area is estimated to be approximately 17.7 inches, based on historical data from eight meteorological stations within a 30 mile radius of the project area. More than half of the annual precipitation occurs during the monsoon season from July through September. The monsoon season is characterized by afternoon thunderstorms that are typically of short duration, but with high-intensity rainfall. The lowest precipitation months are April through June.

Rosemont installed an on-site meteorological station in April, 2006. The station is located at the approximate center of the proposed open pit at an elevation of 5,350 feet amsl. The station monitors site-specific weather data including temperature, precipitation, wind speed, and wind direction. Pan evaporation was added to this station in mid-2008. An annual average pan evaporation of about 71.5 inches was estimated for the Rosemont site based on correlation with data from the University of Arizona and Nogales 6 N weather stations. Pan evaporation rates are approximately 30% higher than open water evaporation due to measurement conditions.

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5.3                     LOCAL RESOURCES

The Rosemont Copper Project is located 30 miles from the city of Tucson with a population of over 520,000 based on the 2010 census. Tucson is also the county seat for Pima County with a population of approximately one million, which comprises the Tucson Metropolitan Area. Mining has been a part of the Tucson area for decades with three major operating mines within a 75 mile radius; including ASARCO Silver Bell Mine near Marana, the ASARCO Mission Complex near Sahuarita, and the Freeport McMoRan Sierrita Mine near Green Valley. There are sufficient resources in the Tucson area for staffing the Rosemont Copper Project. The cultural and educational facilities provided in the Tucson Metropolitan Area will also attract experienced technical staff into the area. There is also a well-established base of contractors and service providers to the mining industry located in Tucson that provide equipment, materials and supplies, as well as maintenance and repair services to the mining industry.

5.4                     INFRASTRUCTURE

The project site is on Arizona State Route 83 (South Sonoita Highway), approximately eleven miles south of Interstate 10 (I-10). This system of state and interstate highways allows quick and convenient access to the site for all major truck deliveries. The majority of the labor and supplies for construction and operations can come from the surrounding areas in Pima, Cochise and Santa Cruz Counties.

The Union Pacific mainline east-west railroad route passes through Tucson, Arizona and generally follows Interstate Highway I-10. The Port of Tucson has rail access from the Union Pacific mainline consisting of a two mile siding complimented by an additional 3,000 foot siding. The siding branches to grade level access, dock level access, intermodal container access, and team track facilities. The Port of Tucson provides transportation and logistics services to businesses in the area and is a registered Foreign Trade Zone. The Port of Tucson is approximately 24 miles from the project site and can trans-load materials and supplies received by rail to trucks for delivery to site.

The Tucson International Airport (TIA) is located approximately 29 miles from the project site and in close proximity to interstate highways I-10 and I-19. TIA provides international air passenger and air freight services to businesses in the area with seven airlines currently providing nonstop service to 15 destinations with connections worldwide. TIA airlines offer over 60 daily departing flights with approximately 6,600 available seats. TIA is designated a US Port of Entry with 24 hour customs and immigration services.

The power supply to the Rosemont mine and production facilities falls within the Tucson Electric Power Company (TEP) and TRICO Electric Cooperative Inc. service territories.

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Geographically, the area east of the Rosemont pit, which includes part of the mine and all of the process facilities, falls in the TEP service territory. The area west of the Rosemont pit, which includes the balance of the mine and the fresh water pumping system, falls within the TRICO service territory. Since most of the estimated electrical load for the project is located in the TEP’s service territory, TEP will be the electrical utility service provider for the entire facility. A joint venture business arrangement will be established between TEP and TRICO to compensate both service providers. Rosemont Copper Company will receive one electric utility rate and bill for the facilities and the breakdown of revenue between TEP and TRICO will be transparent to the project. This multiple service territory and provider agreement will be submitted to the Arizona Corporation Commission (ACC) for review and final approval prior to implementation.

The connection to the TEP power grid will be at the existing TEP South 345/138 kV substation located northwest of Sahuarita, Arizona, in the Santa Cruz Valley west of the project site. A 138 kV power line will be run east and south to a new switchyard (Toro Switchyard) which will be configured with a ring bus to serve existing TEP transmission lines in addition to the new 138kV transmission line to the project site. The new transmission line will follow an alignment along South Santa Rita Road, crossing the Experimental Range, to Helvetia Road at the western base of the Santa Rita Mountains and then it will follow the western access road over the ridge to the plant site.

The most viable source of water supply for the project is from groundwater from various aquifers in the region. Potential sources for groundwater include the basin-fill deposit aquifers of Cienega Wash drainage basin and /or Davidson Canyon located, east and north of the project area, and basin-fill deposit aquifers of the upper Santa Cruz basin west of the project area. There are bedrock and /or shallow alluvium aquifers on or near the Rosemont Project area; however, they are considered to be insufficient as a primary source of water supply for the project. Since the basin-fill deposit aquifers of Cienega Wash drainage basin and Davidson Canyon basin are considered environmentally sensitive, fresh water for the project will be pumped to the project site from new well fields in the basin-fill deposit aquifer of the upper Santa Cruz basin, which lies west of the Rosemont Copper Project and the Santa Rita Mountains. A 53 acre parcel along South Santa Rita Road near the Santa Rita Experimental Range has been purchased and explored with one test well. The test indicated that this property would support 2 wells at a total flow of approximately 3,000 gpm. Water samples collected from the test well indicated that the quality of the ground water is suitable for potable water. It is estimated that a total of 5 or 6 production wells will meet the water supply needs for the project. Additional well sites are under development.

5.5                     PHYSIOGRAPHY

The Rosemont Property is located within the northern portion of the Santa Rita Mountains that form the western edge of the Mexican Highland section of the Basin and Range Physiographic Province of the southwest United States (Wardrop 2005). The Basin and Range physiographic province is characterized by high mountain ranges adjacent to alluvial filled basins. The property occupies flat to mountainous topography in the northeastern and northwestern flanks of the Santa Rita Mountains at a surface elevation ranging from 6,290 to 4,000 feet above sea level.

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Vegetation in the project area reflects the climate with the lower slopes of the Santa Rita Mountains dominated by mesquite and grasses while the higher elevations, receiving greater rainfall, support an open cover of oak, pine, juniper and cypress.

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6                       HISTORY

The early history and production from the Rosemont Property has been described in Anzalone (1995), as well as by Augusta (2007) from which the following summarization is taken.

Sporadic prospecting reportedly began in the middle 1800s in the northwestern portion of the Property and subsequently extended into the eastern part. In 1880, both the Helvetia Mining District (to the west) and the Rosemont Mining District (to the east) were established. Production from mines on both sides of the northern Santa Rita Mountains area supported the construction and operation of the Columbia Smelter at Helvetia on the west side of the range and the Rosemont Smelter in the Rosemont Mining District on the east side of the range. Copper production ceased in 1951 after the production of about 227,300 tons of ore containing 17,290,000 pounds of copper, 1,097,980 pounds of zinc and 180,760 ounces of silver. An unknown, but minor portion of the production came from the Rosemont Deposit.

Since the cessation of production in 1951, the area stretching from Peach-Elgin (on the northwest) to Rosemont (on the southeast) has seen a progression of exploration campaigns to further evaluate the mineral potential. Churn drilling at Peach-Elgin deposit in 1955 and 1956 by Lewisohn Copper Company began the definition of that deposit. Drilling in 1956 by American Exploration and Mining Company initiated exploration of the Broadtop Butte prospect.

By the late 1950s, Banner Mining Co. had acquired most of the claims in the area and had drilled the discovery hole into the Rosemont Deposit. Anaconda Mining Company subsequently acquired the property in 1963 and carried out a major exploration program that demonstrated Rosemont to be a major porphyry/skarn copper deposit, while also advancing regional exploration, including targets at the Broadtop Butte and Peach-Elgin prospects. In 1973, Anaconda joined with Amax to form the Anamax joint venture. The joint venture continued until 1986 when Anamax sold the entire property to a real estate company during the corporate dissolution of Anaconda. By the end of the Anaconda-Anamax programs, exploration drilling totaled about 300,000 feet (91,000 meters), of which approximately 195,000 feet (59,500 meters) define the Rosemont Deposit.

In 1964, Anaconda produced a historical resource estimate for the Peach-Elgin deposit located in the Helvetia District of the northwest part of the Property. Based on assays from 67 churn and diamond drill holes, the historical estimate identified 14 million tons of sulfide material averaging 0.78% copper and 10 million tons of oxide material averaging 0.72% copper. After calculation of that resource, Anaconda and Anamax drilled approximately 140 additional diamond drill holes, but did not update the 1964 estimate. This historical estimate was not prepared to NI 43-101 requirements, but was made by a reputable major copper company and as such is believed to be reasonable as viewed in a historical context. Augusta Resource Corporation has not done the work necessary to verify the classification of this resource and is not treating the resource figure as a NI 43-101 defined resource verified by a Qualified Person and, therefore, the resource figures should not be relied upon by investors.

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In 1977, following years of drilling and evaluation, the Anamax Joint Venture commissioned the mining consulting firm of Pincock, Allen & Holt, Inc. (PAH) to estimate a resource for the Rosemont Deposit. The resulting block modeling calculated a historical geological resource of about 445 million tons of sulfide mineralization at an average grade of 0.54% copper, using a cut-off grade of 0.20% copper. In addition, there were 69 million tons of oxide mineralization at an average grade of 0.45% copper. Subsequent engineering designed a pit based on 40,000 tons per day production rate for a mine life of 20 years. Within the pit design, there were 317 million tons of sulfide mineralization at an average grade of 0.58% copper. In addition, there were 28 million tons of oxide mineralization at an average grade of 0.46% copper. The overall stripping ratio was 3:1 (waste:mineral). The results were described in Pincock, Allen & Holt (1977). This historical estimate was not prepared to NI 43-101 requirements, but was made by a reputable consulting firm and as such is believed to be reasonable as viewed in a historical context. Augusta Resource Corporation has not done the work necessary to verify the classification of this resource and is not treating the resource figure as a NI 43-101 defined resource verified by a Qualified Person and, therefore, the resource figures should not be relied upon by investors.

In 1979, Anamax carried out a resource estimate for the Broadtop Butte deposit located about a mile north of the Rosemont Deposit. Based on the assays from approximately 18 widely spaced diamond drill holes, the historical estimate identified 9 million tons averaging 0.77% copper and 0.037% molybdenum. This historical estimate was not prepared to NI 43-101 requirements, but was made by a reputable major copper company and as such is believed to be reasonable as viewed in a historical context. Augusta Resource Corporation has not done the work necessary to verify the classification of this resource and is not treating the resource figure as a NI 43-101 defined resource verified by a Qualified Person and, therefore, the resource figures should not be relied upon by investors.

ASARCO purchased the property in 1988, renewed exploration of the Peach-Elgin deposit and initiated engineering studies on Rosemont. ASARCO drilling on Rosemont was limited to 12 diamond drill holes.

ASARCO generated a resource estimate of the Rosemont Deposit that was incorporated into a 1997 consulting report by The Winters Company that comprised an “order of magnitude” mining study of the deposit. The resulting “mineable resource,” contained within a pit limit, totaled nearly 295 million tons at an average grade of 0.67% Cu. The plan was based on a production rate of 30,000 tons per day for a mine life of 28 years. The overall stripping ratio was 3.7:1 (waste:mineral). The results and methodology have also been described in Winters (1997). This historical estimate was not prepared to NI 43-101 requirements, but was made by a reputable consulting firm and as such is believed to be reasonable as viewed in a historical context. Augusta Resource Corporation has not done the work necessary to verify the classification of this resource and is not treating the above resource figure as a NI 43-101 defined resource verified by a Qualified Person and, therefore, the resource figures should not be relied upon by investors.

ASARCO sold the entire property to real estate interests in 2004, shortly before the ASARCO takeover by Grupo Mexico S.A. de C.V.

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Augusta Resource Corporation involvement with the Rosemont Deposit began in 2005 and was followed shortly thereafter by an option on the Property, with the completion of the purchase occurring on March 31, 2006. During the option period in 2005, Augusta began a program to confirm the results from previous work and completed a Phase I drilling program consisting of 15 core holes. Based on the new Augusta and previous Anaconda drilling, WLR Consulting, Inc. in conjunction with Mine Reserve Associates, Inc. prepared a mineral resource estimate that was presented in an April 21, 2006 report entitled Mineral Resource Estimate Revised Technical Report For The Rosemont Deposit, Pima County, Arizona, USA. The resource estimate served as the basis for a June 13, 2006 scoping study by Washington Group entitled Preliminary Assessment and Economic Evaluation for the Rosemont Deposit, Pima County, Arizona, USA.

Based on the encouraging results of that program, Augusta continued with a Phase II drilling program in 2006 that consisted of 40 core holes for resource definition, metallurgical, and geotechnical purposes. Additional drill holes were incorporated into a resource estimate update that was announced in a March 16, 2007 press release, which was documented in an April 26, 2007 report, entitled 2007 Mineral Resource Estimate Update for the Rosemont Project, Pima County, Arizona, USA, by WLR Consulting, Inc.

Augusta initiated a Feasibility Study with M3 Engineering & Technology Corporation of Tucson, Arizona in the middle of 2006, which was completed in August 2007. The feasibility incorporates the April 2007 resource model, from which a mineable reserve was established, along the economic evaluation of the overall mine development. The findings were presented in an August 2007 report entitled Rosemont Copper Project Feasibility Study. The development plan presented in the feasibility study was then incorporated into a Mine Plan Of Operation that was submitted to the United States Forest Service, Coronado National Forest, which initiated the National Environmental Policy Act process for permitting surface use of the Forest Service lands.

Augusta conducted further drilling in 2008. Twenty core holes were drilled to further define the northwestern part of the deposit. In addition, ten previously drilled geotechnical holes from Augusta’s 2006 drilling campaign that had not been sampled, were sampled and analyzed. The additional drilling and sampling data was incorporated into a resource estimate that was announced in an October 23, 2008 press release, and was documented by M3 Engineering & Technology Corporation in a January 14, 2009 report entitled Rosemont Copper Project Updated Feasibility Study.

In late 2011/early 2012, Augusta completed a 12-hole, 18,649-foot diamond drilling program, and performed metallurgical testing. Drilling included six holes (7,698 feet) to collect metallurgical test samples, three exploration holes (5,466 feet) drilled to test geophysical targets, and three infill holes (4,711 feet), along with the additional sampling of core remaining from five older holes. The updated drill hole database was used to update the resource model in May 2012 and was the basis for the derivation of updated mineral reserves.

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7                       GEOLOGICAL SETTING AND MINERALIZATION

The regional, local and property geology of the Rosemont deposit is complex and has been studied by numerous geologic investigations including those described by McNew (1981), Anzalone (1995), Hardy (1997), and Augusta (2007).

At Rosemont, Precambrian meta-sedimentary and intrusive rocks form the regional basement beneath a Paleozoic sedimentary sequence of limestone, quartzite, and siltstone. Paleozoic limestone units are the predominant host rocks for the copper mineralization. Structurally overlying these older units at Rosemont are Mesozoic clastic units, including conglomerates, sandstones, and siltstones. Some andesitic volcanic beds occur within the Mesozoic sedimentary section.

The region is characterized by a geologic history that has led to a complex structural character. Compressional tectonism during the Mesozoic and early Cenozoic Laramide Orogeny caused folding and thrust, transverse and reverse faulting, accompanied by extensive calc-alkaline magmatism. Laramide age magmatism, recorded in batholithic and smaller intrusions and their associated volcanic rocks, generated the porphyry copper deposits of the region. At Rosemont, mineralizing quartz monzonite and quartz latite intruded a package of Precambrian intrusive rocks and Paleozoic and Mesozoic sedimentary rocks at the intersection of regional basement structures.

Tertiary extensional tectonism followed the Laramide Orogeny, accompanied by voluminous felsic volcanism. Steeply- to shallowly-dipping normal faults became active during this time, including rotational listric faulting. At Rosemont, it appears that Tertiary faulting has significantly segmented the original deposit, juxtaposing mineralized and unmineralized rocks. The extensional tectonics culminated in the large-scale block faulting that produced the present basin and range geomorphology that is typical throughout southern Arizona.

The generalized regional geology of the Rosemont Property is shown in Figure 7-1. A stratigraphic column of the Rosemont District is presented in Figure 7-2. The local geology of the Rosemont Deposit is shown on a level section in Figure 7-3 and in a vertical section in Figure 7-4.

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Figure 7-1: Regional Geology of the Rosemont Property

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Figure 7-2: Rosemont District Stratigraphic Column

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Figure 7-3: Rosemont Deposit Geology – 4,000 Foot Level Plan

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Figure 7-4: Rosemont Deposit Geologic Section – 11,554,825 E

Drilling to date has defined a significant mineral resource approximately 3,500 feet (1,100 meters) in diameter that extends to a depth of at least 2,000 feet (600 meters) below the surface. Post-mineral features partially delimit the defined resource, dividing the deposit into major structural blocks with contrasting intensities of mineralization. The north-trending, steeply dipping Backbone Fault juxtaposes marginally mineralized Precambrian granodiorite and Lower Paleozoic quartzite and limestone to the west against a block of younger, well-mineralized Paleozoic limestone units to the east. The bulk of the copper sulfide resource is contained in the eastern block of the Backbone Fault. Structurally overlying the sulfide resource is a block of Mesozoic sedimentary and volcanic rocks that contains copper oxide mineralization. These two blocks are separated by the shallowly dipping Low Angle Fault. Other post-mineral features include a deep, gravel-filled Tertiary paleochannel on the south side of the deposit and significant thickness of Cretaceous and Tertiary volcaniclastic material to the northeast of the deposit.

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The bulk of the sulfide resource on the east side of the Backbone Fault and below the Low Angle Fault is hosted in a steeply east-dipping package of Paleozoic-age sedimentary rocks that includes the Escabrosa Limestone, Horquilla Limestone, Earp Formation, Colina Limestone, and Epitaph Formation. The Horquilla Limestone is the most significant, accounting for almost half of the sulfide resource. Significant mineralization also occurs in the Earp Formation and Colina Limestone, as well as in the Epitaph Formation. Relatively minor mineralization occurs in the other Paleozoic units. To the south, the mineralization in this block appears to weaken and eventually die out. To the north, mineralization appears to narrow but continues under cover amid complex faulting. Mineralization continues irregularly to the east of the defined resource, beyond the limit of drilling and beneath an increasingly thick block of Mesozoic sediments.

The Mesozoic rocks of the structural block above the Low Angle Fault consist predominantly of arkosic siltstones, sandstones, and conglomerate. Within the arkose are subordinate andesite flows that range from a few tens of feet to several hundred feet thick. Also structurally wedged into the block at the base of the arkose is the Glance Conglomerate, a limestone-cobble conglomerate, and some occurrences of relatively fresh Concha Limestone.

The Rosemont Deposit copper-molybdenum-silver mineralization is primarily hosted in variable garnet-diopside-wollastonite skarn that formed in the Paleozoic rocks as a result of the intrusion of quartz latite to quartz monzonite porphyry. Marble was developed in the more pure carbonate rocks, while the more siliceous, silty rocks were converted to hornfels. Bornite-chalcopyrite-molybdenite mineralization occurs as veinlets and disseminations in the garnet-diopside-wollastonite skarn and associated marble and hornfels, accompanied by quartz, amphibole, serpentine, magnetite, epidote and chlorite alteration. Quartz latite to quartz monzonite intrusive rocks host strong quartz-sericite-pyrite mineralization with minor chalcopyrite, molybdenite and bornite. Where the mineralized package of Paleozoic rocks and quartz-latite intrusives outcrop on the western side of the deposit, near surface weathering and oxidation has produced disseminated and fracture-controlled copper oxide minerals.

The Mesozoic and lesser Paleozoic rocks above the Low Angle Fault are propylitically altered to an assemblage including epidote, chlorite, calcite, and pyrite. Copper mineralization is irregularly developed. The rocks are commonly deeply weathered and limonitic. The original chalcopyrite is typically oxidized to chrysocolla, copper wad and copper carbonates. Supergene chalcocite is locally present.

Silver occurs in minor, but economically significant quantities in the primary sulfide mineralization in the Paleozoic sequence. The silver is associated with the copper mineralization and is typically tied up in the chalcopyrite and bornite mineral grains. The gold content of the deposit is generally very low, but contributes to a production credit.

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To the north and northwest of the Rosemont deposit are the Broadtop Butte, Copper World and Peach Elgin deposits. These are hosted by intrusive and skarn-altered Paleozoic rocks similar to those at Rosemont, and all are apparently smaller and more structurally- dissected than the Rosemont Deposit.

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8                       DEPOSIT TYPES

The Rosemont Deposit consists of skarn-hosted copper-molybdenum-silver mineralization related to quartz-monzonite porphyry intrusions. Genetically, it is a style of porphyry copper deposit, although intrusive rocks are volumetrically minor within the resource area. The skarns formed as the result of thermal and metasomatic alteration of Paleozoic carbonate and to a lesser extent Mesozoic clastic rocks.

Mineralization is mostly in the form of primary (hypogene) copper-molybdenum-silver sulfides, found in stockwork veinlets and disseminated in the altered host rock. Some oxidized copper mineralization is also present in the upper portion of the deposit. The oxidized mineralization is primarily hosted in Mesozoic rocks, but is also found in Paleozoic rocks where those outcrop or are near-surface on the west side of the Rosemont Deposit. The oxidized mineralization occurs as mixed copper oxide and copper carbonate minerals. Locally, minor amounts of enriched, supergene chalcocite and associated secondary mineralization are found in and beneath the oxidized mineralization.

The Twin Buttes Mine, operated by Anaconda and later by Cyprus, was developed on a deposit with a number of geologic similarities, located about 20 miles (32 kilometers) to the west of Rosemont. The Twin Buttes mine was in production from 1969 to 1994. In addition, the ASARCO Mission Mine, also located about 20 miles (32 kilometers) to the west of Rosemont, has some common geologic characteristics.

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9                       EXPLORATION

Prospecting began in the Rosemont and Helvetia Mining Districts sometime in the middle 1800s and by the 1880s copper production was recorded, which continued sporadically until 1951. By the late 1950s, exploration drilling had resulted in the discovery of the Rosemont Deposit. A succession of major mining companies subsequently conducted exploratory drilling of the Rosemont Deposit and other deposits of the region. Augusta’s work on the deposit has consisted largely of verifying older sampling results, in-fill drilling, and geophysics.

In 2011, Rosemont contracted with Quantec Geoscience to conduct a Titan 24 induced polarization (IP) survey over the Rosemont deposit using the proprietary Titan 24 system, which has the ability to penetrate to depths of several hundred meters. Phase 1 of the survey involved running the system on the known Rosemont deposit to characterize the geophysical response over known mineralization. The geophysical results compared well with the results of the drill hole data from the deposit confirming its usefulness for identifying potentially mineralized targets elsewhere on the property. As a result, Rosemont continued with Phase 2 and Phase 3 geophysical surveys, extending the coverage to the east and north of the Rosemont deposit. A total of 32.4 miles (52.2 kilometers) of geophysical lines were run at Rosemont during the year.

The geophysical surveys identified several anomalies with IP responses similar to that of the Rosemont deposit. One encouraging anomaly was located to the northeast of the Rosemont deposit and was partially drill tested by holes AR-2074 and AR-2080. Drill hole AR-2074 was located where the anomaly reached its shallowest extent, approximately 800 feet (244 meters) below the surface. This hole was drilled to a total depth of 3,500 feet (1,070 meters) and tested the full extent of the IP response, with the top of the anomaly consisting of moderate to strong sulfide mineralization. Starting at 820 feet (250 meters), a 125-foot (38 meter) interval contains 0.82 percent copper, 0.025 percent molybdenum, and 0.19 ounces per ton silver.

A second drill hole, AR-2080, was subsequently drilled to test the IP anomaly further to the south. This hole intercepted low-grade mineralization and alteration in the lower part of the Arkose down to the Low Angle Fault, below which it intercepted minor low-grade skarn in the Epitaph. Deeper in the hole was a significant thickness of graphitic limestone in the Colina Limestone. A minor skarn interval at 1,385 feet consisted of a 15 foot (4.6 meters) thick interval grading 0.95% copper, 0.011% molybdenum, and 0.26 ounces per ton silver. Due to site access circumstances, neither of these holes were drilled directly over the center of the anomaly, and therefore tested only the western edge.

Additional information regarding exploration and evaluations performed on the Rosemont Deposit is presented in Section 6 – History and Section 10 – Drilling.

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10                     DRILLING

Extensive drilling has been conducted at the Rosemont Deposit by several successive property owners. The most recent drilling was by Augusta, with prior drilling campaigns completed by Banner Mining Company, The Anaconda Company, Anamax and ASARCO. Table 10-1 summarizes the drill holes used in the current resource estimate, as they are the holes that are within and adjacent to the deposit. Regional drill holes are archived in a separate data file.

Table 10-1: Rosemont Deposit Drilling Summary

Company Time
Period
Drill Holes
Number Feet Meters
Banner
Anaconda
Anamax
ASARCO
Augusta
1950s-1963
1963-1973
1973-1986
1988-2004
2005-2012
3
113
52
11
87
4,300
136,838
54,350
14,695
132,525
1,311
41,708
16,566
4,479
40,394
Total   266 342,707 104,457

The drill holes utilized in the database were all drilled using diamond drilling (coring) methods. In some cases the tops of the older holes were drilled using a rock bit to set the collar; in other cases the upper parts of older holes were drilled with rotary drilling, switching to core drilling before intercepting mineralization. A map showing the location of the drill holes is provided in Figure 10-1 along with a general outline of the Rosemont deposit limits. Exploration holes drilled using rotary or older “churn” drill holes were excluded from the resource database.

Figure 10-1: Drill Hole Locations in Rosemont Deposit

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In all of the drilling campaigns, efforts were consistently made to obtain representative samples by drilling larger N (1.9 -inch diameter) and H (2.5 -inch diameter) size core. Core recoveries were generally good (typically in the range of 86-93%), lending confidence that quality samples were obtained. Generally, Rosemont drilling is on east-west lines that are approximately 200 feet apart. The average spacing of drill holes along these lines average about 250 feet.

Most of the older Anaconda, Anamax and ASARCO drill core was still available on site or was obtained by Augusta and brought back to the Rosemont Property, where it was systematically re-logged by Augusta personnel to be geologically consistent with the current Augusta drill hole logging. Along with re-logging, this core was also resampled for additional geochemical analyses as described in Section 11 – Sample Preparation, Analyses and Security.

10.1                  BANNER MINING COMPANY DRILLING

The first significant core drilling campaign on the Rosemont Property was by the Banner Mining Company, beginning in about 1961. Banner completed primarily shallow diamond drill holes, many of which were subsequently deepened by Anaconda. Three drill holes included in the resource database were shallow holes started by the Banner Mining Company that were significantly deepened during subsequent Anaconda drilling programs. These holes have a combined length of 4,300 feet.

10.2                  THE ANACONDA COMPANY DRILLING

Anaconda took over Banner’s Rosemont holdings around 1963 and conducted exploration at the Rosemont Deposit and in adjacent mineralized areas. Between the years of 1963 and 1973, they completed 113 diamond drill holes at Rosemont for a total of 136,838 feet. These holes were primarily drilled vertically. Down-hole surveys were conducted during drilling or immediately following drill hole completion for selected holes. Drill hole collars were surveyed by company surveyors. Anaconda drilled approximately 85 percent of the larger N-sized core (1.9 -inch diameter) and 15 percent of the smaller B-sized core (1.4 -inch diameter). Overall core recovery was more than 85 percent.

Exploration subsequently transferred to the Anamax Mining Company (an Anaconda-AMAX joint venture) around 1973, which continued the extensive diamond drilling and analytical work until 1986. Anamax completed 52 core holes for a total of 54,350 feet. These holes were almost exclusively drilled as angle holes inclined -45° to -55° to the west, approximately perpendicular to the east-dipping, Paleozoic, metasedimentary host rocks. Down-hole surveys were conducted during drilling or immediately following drill hole completion for the majority of the holes. Drill hole collars were surveyed by company surveyors. Anamax drilled approximately 80 percent N-sized core (1.9 inch diameter) and 20 percent B-sized core (1.4 inch diameter), with an overall core recovery of more than 88 percent.

During drilling, the core was placed in standard cardboard core boxes by the drillers, with wooden blocks marking the beginning and ending footages of core runs. Core boxes were labeled with the drill hole number, footage interval, and other information by the drillers.

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10.3                  ASARCO MINING COMPANY DRILLING

ASARCO acquired the Rosemont Property in 1988 and conducted exploration until 2004, completing 11 vertical drill holes for a total of 14,695 feet in the deposit area (a 12th hole was drilled to the east of the deposit and is not in the Rosemont Deposit database). Data were available from eight of the ASARCO core holes in the Rosemont Deposit area and were incorporated into Augusta’s resource estimate. Down-hole survey data, if taken, were not available for the ASARCO holes. Drill hole collars were surveyed by company surveyors. The size of core collected by ASARCO was predominantly N-sized (1.9 inch diameter). Core recovery information was not available but Augusta relogging indicated it to be of similar quality to that of other drilling campaigns.

10.4                  AUGUSTA DRILLING

Augusta has conducted diamond drilling in several campaigns, the first starting in the second half of 2005 and continuing into early 2006 (Phase I). The second started in mid-2006 and continued into early 2007 (Phase II). The third started in December 2007 and continued to July 2008 (2008 Drilling). The most recent started in late 2011 and continued into February 2012 (2011/2012 Drilling). In total, Augusta has completed 87 core holes for a total of 132,525 feet (40,394 meters). Of these, 60 drill holes are resource holes to provide infilling of the deposit area, while six were exploration holes outside of the planned pit area, but close enough to be part of the Rosemont deposit database. The remaining 21 Augusta core holes support geotechnical (13) or metallurgical (8) studies. Layne-Christensen, Boart Longyear, and National were the drilling contractors during the Augusta campaigns.

During the 2011-2012 campaign, Augusta drilled 12 holes totaling 18,649 feet (5,684 meters). This included six holes (7,698 feet) drilled to collect metallurgical test samples, 3 exploration holes (5,466 feet) drilled to test a geophysical anomaly, and three infill holes (4,711 feet) drilled in support of a revised resource estimate. Five of the metallurgical test holes and two of the exploration holes were pre-collared with reverse circulation drilling and cased to the top of mineralization. The cored portions of the metallurgical test holes were sampled and assayed for inclusion in the resource database. The infill holes all intercepted significant intervals of copper mineralization and incrementally contributed to the known mineral resources on the northeast edge of the Rosemont deposit. The results of the exploration holes are discussed in the Exploration section of the report. The new drilling was accompanied by the further sampling of five previously drilled holes, the results from which are also incorporated into the new resource model.

Augusta drill holes were usually rock-bitted through overburden, and then drilled with larger HQ-sized core as deeply as possible and finished with NQ-sized core (1.9 -inch diameter) when a reduction in core size was required by ground conditions. As described above, in the most recent drilling, some holes were pre-collared to coring depth. Also in the most recent drilling, several holes were collared with PQ-sized core and then reducing to HQ core once through the incompetent, near-surface oxide zone. This practice significantly improved hole stability and allowed most new holes to be completed with HQ. Augusta drill core was approximately 57 percent N-sized (1.9 inch diameter) and about 42 percent being larger H-sized (2.5 inch diameter), with less than 1 percent being smaller B-sized (1.4 inch diameter). Augusta’s overall core recovery was approximately 95 percent.

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Most of the holes were oriented vertically, although a few of the holes were inclined in order to intercept target blocks from reasonably accessible drill locations. All drill holes were surveyed down-hole with a Reflex EZ-Shot survey instrument that measured inclination/dip and azimuth direction, with readings generally taken every 100 feet down the hole during 2008 and every 200 or 500 feet down the hole during 2005, 2006 and 2011-2012. Phase I drill hole collar locations were surveyed by Putt Surveying of Tucson, Arizona, while all later drilling locations were measured by Darling Environmental & Surveying.

During drilling, the core was placed in standard cardboard core boxes by the drillers, with wooden blocks that marked the footages of core runs. Core boxes were labeled with the drill hole number, footage range and other information by the drillers.

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11                     SAMPLE PREPARATION, ANALYSES AND SECURITY

The Rosemont resource database is based on core samples recovered from diamond drill holes. The drill core from mineralized intervals was generally sampled continuously down the hole, at a nominal five-foot sample length. In taking a sample, the core is generally halved (split) along the long axis, taking care to evenly distribute veinlets and other small-scale mineralized features where present, into both halves of the core.

11.1                  BANNER, ANACONDA AND ANAMAX SAMPLING AND ANALYSES

The Banner, Anaconda and Anamax sampling are discussed as a group because the sampling took place as part of a more-or-less continuous program. The analytical data in the resource database for the three Banner drill holes came from the Anaconda laboratory, as most of the length of these holes came from subsequent Anaconda drilling that significantly deepened these holes. The exploration transition from Anaconda to Anamax (Anaconda-Amax Joint Venture) drilling did not immediately utilize a different laboratory or techniques.

The core was first logged to record the core run intervals and percent recovery, along with lithology, structure, alteration and mineralization. After sampling intervals were assigned, the core was split with a mechanical splitter along its long axis, and one-half of the core was retained in the original core box. Sample preparation during the Banner, Anaconda and Anamax programs was conducted by employees of those companies. Other details of the sampling process are not well known, but since this work was carried out by major copper companies for their internal use, it is believed that they used the standard industry practices for that time.

The core was sampled at geologic intervals, based on changes in mineralization and alteration. Intervals were from one to six feet in length and averaged about five feet. In poorly mineralized intervals, analytical samples were collected only intermittently, typically with one five-foot sample collected every 20 to 30 feet, to characterize the rock as having low to no grade values.

The Banner, Anaconda and Anamax geochemical suite was determined by whether an interval retained its primary sulfide mineralization or had been oxidized. Core with primary sulfide mineralization above trace levels was comprehensively analyzed for total copper and molybdenum. For some intervals, lead and zinc metal concentrations were analyzed where indicated by mineralogy, but that was not common. Relatively late in the program, particularly in the Anamax drill core, silver analysis was routinely included in the sulfide zone, especially for well-mineralized intervals. Oxide zone drill core with visible copper oxide mineralization (chrysocolla, cuprite, copper wad, etc.) was analyzed for acid-soluble copper in addition to total copper, while molybdenum was excluded or only intermittently analyzed in the oxide zone core.

Details of the analytical methods used at the Anaconda and Anamax laboratories were outlined by Mr. Dale Wood, Anaconda Chief Chemist in meetings and telephone conversations on November 28, 2005 and January 21, 2006. Crushing and grinding reduced all pulp samples to minus 100 mesh size, with constant screen size testing. Copper and molybdenum were determined by x-ray fluorescence (XRF) screening and then wet chemical methods, using analytical procedures that were industry standard for the 1960s and 1970s. Samples with XRF-determined grades above 0.2% Cu and 0.02% Mo were selected for wet chemical analyses. Pulp samples for the wet chemical method were brought into solution by hot acid digestion on a shaker table with hydrochloric acid, nitric acid and perchlorate acid added to the boiling solution followed by a few drops of hydrofluoric acid. Analyses for molybdenum were by the colorimetric iodine titration method. Copper analyses were done by the colorimetric phenolthylanaline titration method. The XRF analytical technique consisted of either a quick screening method by compressing a pulp sample on mylar film and placing it under the x-ray beam or, alternatively, adding cellulite to the pulp sample, pressing it into a ring and then placing under the x-ray beam.

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11.2                  ASARCO SAMPLING AND ANALYSES

The ASARCO drill core was routinely analyzed for total copper, acid-soluble copper and molybdenum. Oxide zone core does not appear to have been analyzed differently than the sulfide zone core. The core was sampled with preference towards a 10-foot sample length, but longer or shorter intervals were sometimes used. The ASARCO drill core was apparently logged and sampled in much the same style as is described above for the Banner, Anaconda and Anamax core.

The ASARCO geochemical analyses that Augusta obtained from ASARCO were conducted by Skyline Analytical Laboratory, Tucson, Arizona. Skyline is a large, certified, commercial laboratory that utilized industry-standard analytical techniques and therefore the data obtained for the ASARCO core are considered reliable. No detailed descriptions of Skyline’s sample preparation and analytical methods during those years are available.

11.3                  AUGUSTA SAMPLING AND ANALYSES

11.3.1               Augusta Core

For the 2005 Phase I and the 2006 Phase II drilling programs, sampling of Augusta drill holes took place at the Rosemont Ranch sampling facility. The 2008 and later drill hole sampling took place at the Hidden Valley Ranch sampling facility. Geotechnical logging was performed on all core drilled by Augusta to systematically quantify Rock Quality Designation (RQD), core recovery, fracture frequency, core hardness, joint condition and large-scale joint expression. Core logging geologists familiar with the project recorded the rock type, alteration, mineralization, and structure. After logging, the geologist assigned and marked the sample intervals and cut-lines directly on the core and on the core box interior with a black marker. Each sample was given a unique, sequenced sample number with the footage noted in a sample tag booklet and in Excel-based spreadsheets. The drill core boxes were then photographed with a digital camera.

Augusta core was sampled at even five-foot intervals, except where massive copper or molybdenum veining, structures or lithologic breaks warranted special investigation through the selection of shorter intervals. Sample intervals would return to footages evenly divisible by five as soon as possible thereafter. This tended to occur in earlier campaigns and was not a practice during more recent campaigns. One exception is the sampling from the 2012 metallurgical holes, which were sampled at 10-foot intervals.

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The core was split by cutting it in half with a diamond rock saw. All cuts were carefully planned and marked on the core by the logging geologist to evenly divide mineralization between the two halves of the core. All core cutting was done with water using no additives and the sawed drill core was placed directly back in the core box to dry before sampling. When dried, the left-hand half of the split core was placed in bags labeled according to the sequenced paper sample tags, with a sample tag also placed inside the bag. The plastic bags were then sealed with plastic ties, leaving the sample number visible.

Core was analyzed using a geochemical suite that varied depending on whether or not the core retained its primary sulfide mineralization or had been oxidized, similar to the approach described above for Anaconda. In the oxidized zone, the core was routinely analyzed for total and acid-soluble copper. Sulfide zone core was analyzed for total copper, molybdenum and silver. In 2005 and 2006, some core was also analyzed for gold, although that was discontinued when the gold content had been adequately characterized and the cost of additional gold analyses was no longer warranted.

Geochemical analyses for Augusta-drilled core and for the Augusta resampling of the Anaconda, Anamax, and ASARCO core were primarily performed by Skyline Assayers and Laboratories (Skyline) in Tucson, Arizona. During 2005, Skyline was formally known as Actlabs-Skyline and had been owned by ACTLABS (Ancaster, ON, Canada) since 1997. Skyline became independent of ACTLABS in January, 2006. Skyline is accredited in international quality standards through ISO/IEC 17025, with CAN-P-1579 for specific registered tests through the Standards Council of Canada. Skyline is considered to be a reputable and trustworthy facility and is used by a number of major and junior mining companies in the southwest area of the United States.

Augusta had both primary and secondary (duplicates) analyses done at Skyline in 2006 and 2007. ALS Chemex (Vancouver, BC, Canada) analyzed duplicate checks samples in 2005. ALS Chemex has accreditation through ISO 9001:2000 in North America.

At Skyline, the entire sample was crushed using a TM Terminator to produce a greater than 80% pass 10-mesh product. Samples were blended and divided using a two-stage riffle splitter, from which a 300-400 gram split was pulverized to a 90% passing 150-mesh product using a TM Max 2 Pulverizer. Wash gravel and sand were used by Skyline to clean the crushers after each batch of samples were processed. Pulverizers were cleaned after each batch of samples and/or after each sample if the material adhered to inside walls of the grinding vessel. Coarse reject and pulp material was saved and returned to Augusta.

For the determination of total copper and molybdenum, Skyline digests 0.2000 to 0.2300 grams of the sample with 10.0 milliliters (ml) of hydrochloric acid, 3.0 ml nitric acid and 1.0 ml perchloric acid at 250° C, in a 200-ml phosphoric acid flask. When the only remaining acid present is perchloric acid and the volume of the liquid in the flask is less than 1 ml, the solution is allowed to cool. About 25 ml demineralized water and 10.0 ml hydrochloric acid is then added and the solution is gently boiled for 10-20 minutes. The flask is again cooled to room temperature and the contents are diluted with demineralized water and shaken well to mix. Copper is determined by atomic absorption. Molybdenum is determined by ICP.

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Acid soluble copper is determined by leaching one gram of pulverized sample in 10% sulfuric acid solution for one hour at room temperature. The copper content of the resulting solution is determined by atomic absorption.

For the determination of silver, Skyline digests 0.25 grams of sample with 0.5 ml nitric acid and 1.5 ml hydrochloric acid in a disposable, 18-mm x 150-mm borosilicate glass test tube. After agitation and the cessation of any effervescence due to carbonates, the test tubes are placed in a test tube rack in a hot water bath that is maintained between 90 °C and 95 °C, where digestion continues for 90 minutes. After cooling to room temperature the contents are diluted to 10 ml with demineralized water and again agitated to mix well. The solutions are then read by atomic absorption for silver.

11.3.2               Banner, Anaconda, Anamax, and ASARCO Core Resampling and Analysis

Augusta extensively resampled core drilled by Anaconda, Anamax, and ASARCO, most of which was available, to fill-in missing analytical information and to validate the older analyses. Resampling of pre-Augusta drill holes took place at the Hidden Valley Ranch sampling facility in 2006. Augusta geologists identified intervals requiring additional (infill) analyses by referring to the previous logging and analyses for the core. New Augusta sample intervals were assigned unique, sequenced sample numbers from sample tag books in which hole identification and interval footage were recorded, and this information was recorded in an Excel-based spreadsheets. The core boxes were carefully photographed using a digital camera, and the photos were inspected and archived before samples were collected. The assigned intervals were measured and collected by sampling technicians, taking the entire remaining core with the exception of some small, representative archive samples. The individual samples were placed in plastic sample bags marked with the new sample number and the paper tags from the sample books were placed in the sample bags before the bags were sealed with plastic ties.

Whenever possible, the sample intervals for additional analyses conformed to pre-existing sample intervals, allowing the old and new data to be easily combined and compared. Augusta required all samples to be seven feet or shorter. If only intermittent samples had previously been collected (i.e., a five-foot sample every 20-30 feet), unsampled original intervals were divided into multiple new sample intervals of approximately five feet in length, preserving the starting and ending footages of the original sample intervals. If core was missing, either lost or previously taken for metallurgical work, Augusta sample intervals were aligned to reflect the missing core intervals.

Oxide zone intervals were analyzed for both total and acid-soluble copper if total copper was estimated to be >0.1% Cu, but acid soluble copper data were not available. All sulfide zone drill core from within the deposit area that had not been analyzed for both total copper and molybdenum was sampled and analyzed to provide complete, continuous copper and molybdenum data.

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Silver and to a minor extent gold were analyzed for drill hole intervals that were missing these values in the historic data. Sulfide zones for which previous copper analyses indicated an interval contained greater than 0.2% Cu was sampled by means of a composite representing a 50-foot continuous interval length. Gold analyses were discontinued late in 2006 after the gold mineralization was sufficiently characterized. For the purposes of silver and gold analyses, the composite sample intervals were combined into length-weighted 50-foot sample before analysis, thereby reducing the total number of samples. This compositing was performed on pulp samples at the analytical laboratory using relative weight contributions for each component sample calculated by Augusta geologists.

The analytical procedure for the core resampling program were the same as described above for the Augusta drill holes.

11.4                  SAMPLE HANDLING AND SECURITY

Sample handling during the historic Banner, Anaconda, Anamax, and ASARCO programs was conducted by employees of those companies, for which some of the protocol records are limited. Augusta notes that these were major mining companies conducting work for their internal use. It is assumed that professional care was taken in the handling of samples by these company employees and no evidence to the contrary has been found.

For the new Augusta drilling program, the drilling contractors kept the core in a secure area next to the drill rig before delivering it to the Rosemont Ranch (2005, 2006) or Hidden Valley (2008-2012) sampling facility, approximately three miles from the drilling area. At the Rosemont Ranch facility in 2005 and 2006 and subsequently at Hidden Valley in 2008, samples were logged, marked, cut and placed in sample bags by geologists and helpers contracted by Augusta. At both locations, for programs through 2008, the samples were kept in locked storage units on site until they could be transported to the analytical laboratory in Tucson. The logging and sampling areas were kept under closed-circuit video surveillance to provide a record of the personnel that had accessed the logging and sampling areas. Additional security was afforded by ranch personnel that oversaw the premises at night. For the 2011-2012 drilling, the locked storage units and video surveillance were superseded by 24 hour-per-day private security guards. No core handling or core security issues were experienced during the drilling or sampling programs.

Locked sample boxes were picked up by Skyline employees, who officially took custody of the samples at the two sampling facilities, set up on the Rosemont Property. After completion of the laboratory work, the pulp samples and coarse rejects were returned to site for long-term storage and possible future use.

11.5

QUALITY ASSURANCE AND QUALITY CONTROL

   
11.5.1

Historic Protocols

The Quality Assurance and Quality Control (QA/QC) protocols in place during the Anaconda, Anamax and ASARCO exploration programs are not documented in records available to Augusta, although all the available evidence shows that they took great care in sample handling and storage, and that the laboratories analyzing the geochemical samples used industry standard practices.

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11.5.2               Augusta Protocols

Rosemont verified the accuracy and precision of its geochemical analyses by inserting standards of known metal content in the sample stream at periodic intervals and by reanalyzing approximately 5% of all samples to check the repeatability of results. Rosemont’s QA/QC protocol was provided initially by Kenneth A. Lovstrom (deceased), Geochemist. After January 2006, the protocol was subsequently directed by Shea Clark Smith, Geochemist and Principal of Minerals Exploration & Environmental Geochemistry. Details of Rosemont’s QA/QC procedure are as follows:

1.

Standards were submitted with a frequency of one per 20 samples. The inserted standards were chosen to be similar in grade to the drill holes samples that they accompanied whenever possible.

2.

Blank samples were submitted with a frequency of one per 40 samples.

3.

Approximately 5% of all samples were reanalyzed in what was called the Check Assay Program.

As part of the protocol, whenever standards or blanks returned from the laboratory with values significantly different from what was expected, the standard or blank pulp was resubmitted to the laboratory along with two samples that occurred on either side of the questionable standard or blank in the sample stream. In most cases this process validated the initial analyses. If not, the entire job was rerun, which only occurred in a couple of rare instances since 2005.

In addition to Rosemont’s standards and repeat analyses, further QA/QC was provided by the results from other standards inserted into the sample stream by the assay lab, Skyline Assayers and Laboratories, Tucson, Arizona (Skyline). The results from those standards are reported on Rosemont’s assay certificates.

11.5.3               External Augusta Standard Reference Materials

Since 2005, Rosemont has used 14 standard reference materials (SRM) incorporating a range of copper, molybdenum, and silver concentrations that approximate the range of metal values encountered in Rosemont’s analytical samples. The SRM used in 2005 were KM-5, GRS-3, GRS-4, OC-43, OC-48, R1, and R2. Of these, KM-5, GRS-3, GRS-4, OC-43, OC-48 were developed by Mr. Lovstrom and the detailed analytical results on which their Certified Values are based are no longer available (they were only used minimally).

A new suite of SRM was created specifically for the Rosemont Deposit in 2005 and 2006 and includes: R1, R2, R4A, R4B, R4C, R4D, R4E, R4F, and R4G. These were prepared at MEG Labs (Carson City, NV) from naturally mineralized rock that had been collected at the Rosemont Deposit. The metal values for these standards were established by a round robin analytical program, compiled from a minimum of 25 samples of each SRM that had been sent to 5 or more laboratories. The average values and standard deviations calculated from the round robin program establish MEG Labs Certified Values for the R1, R2, and R4 suite of standards. It is noted that there is a good balance between the known copper grades of the R-series SRMs and the average economic metal concentration in the drill samples being run with these SRMs.

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For Rosemont samples, the performance of mineralized SRM in the analytical stream is good to excellent for copper and molybdenum. Figure 11-1 shows the copper results for four of the more widely used external standards. Table 11-1 summarizes the performance of each of the standards used during the various sampling campaigns. It is noted that for little-used (2005 only/limited number of samples) Lovstrom standards GRS-3 and GRS-4, the reported total copper concentrations are up to ten percent lower than expected for copper. The issues with these SRMs were quickly identified and their use was terminated.

For the most widely-used standards, copper values returned are typically within about +/-0.01% Cu of the expected value. No significant, systematic bias in copper values is apparent across the range of results. These results show that copper, by far the most important metal in the deposit, has reliable assays.

Molybdenum results are good, with molybdenum from mineralized SRM generally reported within +/- 0.002% Mo of the established value for most standards used. In the low-grade range of molybdenum commonly found in the deposit, a relatively larger percent variability is commonly observed. Generally, the analytical laboratory is reporting acceptable levels of molybdenum, with a slight tendency to report lower than the expected SRM concentrations of molybdenum.

Silver results are fair to good, with silver from mineralized SRMs generally reported within +/-0.015 opt Ag of the established value for most standards used. The results for silver, a minor economic constituent of the deposit, are influenced by the low-grade range typically being considered. In this low-grade range, the relative percent variability is higher, particularly as values get closer to the analytical detection level. Silver concentrations from the analytical laboratory are within acceptable levels of silver, but there is a tendency to sometimes report lower than expected SRM concentrations. The analysis of Ag at the level of several parts per million or less is inherently difficult, affected by issues such as sample-dependent correction requirements, sample homogeneity, and reliability of sample digestion.

The performance of the SRMs in the analytical stream was acceptable for the three economic metals under consideration.

11.5.4               Internal Skyline Laboratory Standard Reference Materials

Internal standards used by Skyline indicate that accuracy is within tight tolerances of 0.00% to 0.03% for copper, 0.000% to 0.001% for molybdenum, and 0.000 to 0.015 opt for silver. Copper shows no systematic bias.

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Figure 11-1: QA/QC Standard and Blank Performance

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Table 11-1: Rosemont and Skyline Standards Qa/Qc Results


Stan ID

N4
Cu (%) Mo (%) Ag (opt)
Expected Found % Var  Notes2  Expected   Found % Var Notes2 Expected   Found % Var Notes2
Rosemont Standards                               
R1 462 0.47 0.47 0 1 rej        0.025 0.025 0 6 rej 0.149 0.149 0 6 rej
R2 278 0.72 0.71 -1          0.017 0.018 6   0.207 0.204 -1  
R4A 2 1.43 1.40 -2          0.032 0.031 -3          
R4B 54 0.57 0.57 0 1 rej        0.030 0.029 -3 1 rej 0.114 0.110 -3 1 rej
R4C 133 0.39 0.40 3          0.033 0.031 -6   0.090 0.072 -20  
R4D 89 0.30 0.30 0          0.018 0.018 0   0.069 0.059 -14  
R4E 135 0.22 0.21 -5 1 rej        0.011 0.009 -18   0.051 0.037 -27 1 rej
R4F 106 0.14 0.14 0 3 rej        0.010 0.009 -10   0.041 0.029 -30  
R4G 22 0.07 0.07 0          0.016 0.014 -13   0.035 0.016 -54  
Blank 792 27/655 >dl5, those average 0.01%  74/536 >dl, those average 0.002% 269/508 >dl, those average 0.010 opt
Lovstrom Standards3      
KM-5 90 0.99 1.01 2                  
GRS-3 19 1.23 1.12 -9                  
GRS-4 18 2.02 1.9 -6                  
OC-43 25            0.035 0.034 -3          
OC-48 21            0.078 0.074 -5          
Skyline Standards1      
CGS-2 255 1.18 1.16 -2 1 rej                
CGS-3 271 0.65 0.65 0 1 rej                
CGS-4 184 1.95 1.93 -1 1 rej                
CGS-6 216 0.32 0.32 0 1 rej                
GXR-1 121                 0.905 0.867 -4  
GXR-2 121                 0.496 0.481 -3  
GXR-4 118                 0.117 0.106 -9  
HV-2 295 0.57 0.58 2          0.048 0.047 -2          
CGS-25 25 2.19 2.16 -1                  
CM-1 53 0.85 0.85 0          0.076 0.076 0          
CM-2 33 1.01 1.00 -1          0.029 0.028 -3          
CM-8 52 0.36 0.37 2          0.016 0.016 0          
Cu-121 15 0.97              0.042       0.964 0.975 1  
Cu-122 34 0.79 0.77 -3          0.076 0.077 1   2.132 2.143 1  
Cu-123 78 0.49              0.051       1.256 1.275 2  

1. These standards are run by Skyline as part of their ongoing Qa/Qc program and are reported at the request of the client. Standards GXR-1, 2, and 4 are only pertain to Ag composite samples from the core reassay program (2006-2007), whereas the other Skyline standards reflect a broader scope of analytical work.

2. Reject (“rej”) samples are those that are removed from the population of standards for the purposes of average calculations and charting. They are interpreted to have been physically-switched or misidentified by either Rosemont or Skyline personnel.

3. Lovstrom standards were used in 2005 only. They were provided by Ken Lovstrom, geochemist (deceased). Documentation of their metals content is lacking.

4. N is the number of times a standard was used in the Qa/Qc program.

5. dl is the detection limit for the analytical analysis.

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11.5.5               Blank Samples

Blanks, made of barren quartz sand, do not contain metal concentrations above the limit of detection. These were submitted with drill cuttings with a frequency of one per 40 drill samples. Blanks test the laboratory performance at the limit of detection and can reveal problems with contamination between samples.

Out of 655 copper assays reported for the blanks, only 27 (4%) reported concentrations for copper above the limit of detection (0.01% Cu), and 22 of those 27 values were at the threshold limit of 0.01% Cu. The highest copper value reported for a blank was 0.07% Cu.

For molybdenum, 74 of 536 (14.0%) values reported for the blank were above the limit of detection, and 64 of the 74 values were at the threshold value of 0.001% Mo. The highest value reported for a blank was 0.013% Mo, and the average of non-blank values was 0.002% Mo.

For silver, there were 269 out of 508 (53%) values reported for the blank that were above the limit of detection, of which 77 values were at the threshold value of 0.003 opt Ag (0.1 ppm Ag). The average concentration of those samples that reported above the detection limit is about 0.010 opt Ag (0.34 ppm). These silver results further demonstrate the variability of very low-level (non-economic) Ag analyses. It is noted that the performance of the blanks at Skyline has improved with Rosemont QA/QC discussions that have allowed for improvement to the analytical procedures.

The performance of the blanks in the analytical stream was acceptable for the three economic metals under consideration. The incidence of blanks reporting metals values was very minor, with the exception of silver, where economically insignificant low-grade values were sometimes reported.

11.5.6               Check Assays (Pulp Rerun Analysis)

Approximately 5% of the samples (new drilling and resampling by Augusta) were resubmitted to Skyline Assayers and Laboratories of Tucson, Arizona, at the end of each drilling campaign in what was called the Check Assay Program. Samples consisted of the originally prepared pulp material that was resubmitted for analysis. All samples analyzed were total copper, acid-soluble copper, molybdenum and silver.

A suite of standards described below accompanied each analytical batch in the laboratory during the Check Assay Program. In each batch there are 16 core samples for reanalysis, accompanied by up to four Augusta standards and as many as five Skyline internal standards. In addition to the inclusion of standards, further QA/QC validation of the check assays is provided by Skyline’s reporting of repeat AA readings for the first and last of Augusta’s 20 samples and standards in each analytical batch.

Generally, results for total copper compare quite precisely and with no significant bias from values being systematically higher or lower than original values. Considering samples with copper contents greater than the detection limit, 91 percent of the samples have comparison assays that are within 10% of each other. This provides good confidence in the repeatability of the copper analyses. The summarized results are presented in the attached Check Assay Summary Table shown below. The check assay results are summarized in Figure 11-2 and in Table 11-2.

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Molybdenum results show somewhat greater deviation (lower precision) with 61 percent of the samples returning comparison assays within 10% of each other. Again, there is no significant bias indicated by the average difference between the original and check assay. It is noted that the molybdenum grades being analyzed are very low and the associated broader variability observed is typical in these lower grade ranges.

Silver comparisons indicate less precision, with 44 percent of the samples returning comparison assays within 10% of each other. As in the other elements, there is no significant bias for the check value to be higher or lower than the original values. It is noted that the silver grades being analyzed are very low and approach threshold limits, were broader variability would be expected. A number of silver checks are different enough to indicate possible sample switches. That issue has not introduced a bias in data.

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Figure 11-2: Check Assay Performance

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Table 11-2: Check Assay Performance Summary

Augusta Resource Rosemont Project
Check Assay Summary Table1

Copper Check Assay Summary          
Cu > Detection
Limit (N=572)
% Variance2 <5 <10 <25 <100
n 462 523 552 572
% of N 81 91 97 100
Original Assay-Check Assay: Average Value = 0.00% Cu

Molybdenum Check Assay Summary          
Mo > Detection
Limit (N=429)
% Variance2 <5 <10 <25 <100
n 191 260 350 429
% of N 45 61 82 100
Original Assay-Check Assay: Average Value =0.000% Mo

Silver Check Assay Summary          
Ag > Detection
Limit (N=732)
% Variance2 <5 <10 <25 <100
n 188 321 525 732
% of N 26 44 72 100
Original Assay-Check Assay: Average Value = -0.1 opt Ag

Notes:
1. 5% of 2005-2006 samples from Augusta's AR drill holes and the Anaconda/
Anamax/Asarco drill hole resample program were submitted for check analyses.
2. % Variance = Absolute value of 100*((Origninal Assay-Check Assay)/Original Assay);
% variance is always positive in this table.

11.6                  QA/QC SUMMARY

The analytical QA/QC program demonstrated that the copper, representing approximately 80% of the deposit value, is well behaved based on sample QA/QC work. Molybdenum and silver, accounting for approximately 15 percent and 5 percent of the deposit value, respectively, experience a little more variability, which can largely be attributed to the low concentration levels of these metals in the samples that are analyzed. Overall, all three metals are considered reliable for resource estimation work. The most recent drilling program results are similar to the overall QA/QC results obtained since 2005 by the various Augusta sampling campaigns.

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12                     DATA VERIFICATION

Augusta took a number of steps to verify the results of earlier exploration results by other companies. These previous efforts were conducted by recognized major companies and it is believed their work was conducted to industry standards at that time. Augusta’s own work was conducted with appropriate sampling handling and QA/QC measures to ensure that resulting data were reliable. Quality control measures for sample assaying are described in detail in Section 11.

12.1                  TEN HOLE RESAMPLING PROGRAM

Anaconda, and to a lesser extent Anamax, and to a minor extent ASARCO generated a portion of the geochemical data in the resource database. Augusta performed significant resampling and assaying of the older drill holes to fill in missing data, but typically did not generate replacement data. In order to directly validate the old data with comparable values from Skyline Laboratories, Augusta reanalyzed 10 historic drill holes (5 Anaconda (A-xxx), 4 Anamax (xxxx) and 1 ASARCO (AH-x)) in their entirety. The remaining ½ split of core material from the 10 historic holes was collected in sample intervals corresponding to the original sample intervals and assayed for Cu, Mo and Ag. For silver, several of the historic holes did not have previous analyses to compare with the new values. The results are tabulated in Table 12-1.

Table 12-1: Ten Hole Resampling Program Summary

  Cu (%) Mo (%) Ag (opt)
  Old Hole Augusta Old Hole Augusta Old Hole Augusta
All Hole
Average
0.50 0.48 0.022 0.017 0.31 0.20
Individual Drill Hole Comparisons
A-804 0.45 0.43 0.007 0.005 --- ---
A-813 0.51 0.52 0.033 0.028 --- ---
A-821 0.57 0.53 0.038 0.029 --- ---
A-834 0.48 0.45 0.018 0.018 --- ---
A-858 0.35 0.34 0.018 0.016 --- ---
1485 0.43 0.38 0.017 0.005    
1508 0.94 0.90 0.024 0.023 0.31 0.23
1916 0.39 0.39 0.026 0.015 0.31 0.17
1917 0.23 0.25 0.023 0.012 0.30 0.13
AH-4 0.37 0.38 0.009 0.009 --- ---

Generally, new (Rosemont) values for total copper are quite similar to the old (Anaconda, Anamax, ASARCO) values. Overall, the averages for all comparison intervals are 0.50% Cu (old) vs. 0.48% Cu (new). It is believed that the small amount of difference is due to variability in the distribution of mineral grains in the core.

Molybdenum results show more variability between old and new values than do the copper results, with a tendency for old molybdenum values to be higher than new molybdenum values. The difference is somewhat attributable to the presence in the database of old (1964-1983) low-grade, x-ray analyses from Anaconda and Anamax laboratories. Typically, samples were screened with the x-ray technique, and only run with the more accurate wet chemical technique if the x-ray results indicated a value greater than 0.020% Mo or 0.20% Cu.

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A statistical study was conducted by MRA (2006) on all available data to determine the correlation coefficient between XRF and wet chemical values for both Cu and Mo. The study showed excellent agreement with correlation coefficients of 0.944 for Cu and 0.874 for Mo. MRA concluded that these results indicate that the lower grade XRF values would be valid for use in grade estimation in the model for both Cu and Mo. It is noted that the XRF values would now usually only remain in the database for lower grade intervals of limited relevance to the project viability.

Silver results also show greater variation between old and new values than do the copper results, although only three holes can be compared because of the lack of historical silver values for these particular holes. It is noted that the combination of Augusta’s own drilling and Augusta’s resampling of older core where silver was missing, has resulted in a majority of the silver data in the current resource database to be the newer Skyline analyses.

12.2                  ADJACENT (METALLURGICAL) COMPARISON HOLES

To further compare historic data to newer Augusta data, the results from metallurgical holes drilled by Augusta in 2011 were compared to adjacent historic holes. The holes were not necessarily twins, as they had separation distances that ranged from 13 to 39 feet. The metallurgical holes were assayed as are typical resource drill holes, and provide 6 pairs of adjacent drill holes for comparison of the metal contents of the historic holes. Table 12-2 compares the adjacent hole metal contents.

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Table 12-2: Adjacent Drill Hole Comparison Summary

Drill Holes Compared Interval Thickness Cu (%) Mo (%) Ag (opt)
Average Met Hole Value:     0.656 0.015 0.19
Average Historic Hole Value:     0.633 0.018 0.19
AR-2075 (met hole) 490'-1590' 1100' 0.63 0.017 0.20
A-831 507'-1592' 1085' 0.55 0.019 0.17*
(25' collar spacing)          
AR-2076 (met hole) 320'-1100' 780' 0.48 0.016 0.18
A-809 305.5'-1126' 820.5' 0.55 0.012* 0.26*
(17' collar spacing)          
AR-2078 (met hole) 720'-1350' 630' 0.52 0.011 0.12
A-841 736'-1350' 614' 0.65 0.020 0.12
(18' collar spacing)          
AR-2079 (met hole) 740'-1490' 750' 0.99 0.013 0.22
A-823 740'-1491' 751' 0.71 0.025 0.17*
(39' collar spacing)          
AR-2082 (met hole) 228-810 582' 0.57 0.011 0.24
A-846 228-813 585' 0.45 0.016 0.18*
(13' collar spacing)          
AR-2077 (met hole) 810'-1320' 510' 0.75 0.020 0.20
AR-2010 818'-1323' 505' 0.89 0.018 0.27
(22' collar spacing)          
* historic data incomplete

The average copper grade of all metallurgical drill holes is 0.656% Cu, compared to 0.633% Cu for the average of all adjacent drill holes. For most comparison pairs, copper values show only minor variability. It is believed that the small amount of difference is due to natural geologic variability that exists over the 13 to 39-foot spacing between the holes. No bias is indicated, as there was no systematic difference between the older hole and the Augusta hole when comparing the grade pairs.

The average molybdenum grade of all metallurgical holes is 0.015% Mo, compared to 0.018% Mo for the adjacent drill holes. For most comparison pairs, molybdenum values show moderate variability, most of which can be attributed to natural geologic variability between the holes. There is also no bias indicated when comparing the grade pairs.

Average silver values show negligible overall differences in the overall averaged values for all comparison pairs. Both metallurgical holes and adjacent holes contain 0.19 opt Ag on average. Individual pairs, however, show moderate variability, some of which is attributable to the low silver levels being analyzed for and some being natural geologic variability. Comparison of the grade pairs does not indicate a bias.

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This comparison of the metallurgical drill holes with adjacent drill holes contributes to the validation of the three metals of economic interest. Overall, the grades are comparable and in all cases are considered to reasonably reflect the grades of the Rosemont deposit.

12.3                  DRILL HOLE DATA ENTRY VALIDATION

In order to assess the integrity of the data entry in the drill hole database, the database has been inspected by MRA after the completion the 2005, 2006 and 2008 drilling programs. MMTS inspected the more recent data entry related to the 2011 drilling. A visual inspection was conducted comparing a random sampling of the values shown on the original assay certificates to those listed in the database files to check for data entry errors. The number of data errors found was minimal from all of these data entry checks and some required relatively insignificant changes. From this it was concluded that the data entry into the drill hole database was reliable for use in the resource modeling.

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13                     MINERAL PROCESSING AND METALLURGICAL TESTING

The earliest existing records of metallurgical testing are from the period 1974 - 1975, at which time grinding and flotation tests were performed. In the first half of 2006, Augusta initiated test work to provide a better understanding of the metallurgy of the Rosemont Deposit and establish the design criteria for the design of a process facility.

The samples tested both as deposit composites and individual variability samples and are considered to fairly represent the deposit. Details of the individual samples used to make up composites are described in the individual test reports.

The copper sulfide ore contains two main types of copper mineralization: chalcopyrite and bornite/chalcocite. There are three major and several minor lithological units within which the two types of sulfide mineralization occur:

  • Horquilla
  • Earp
  • Colina
  • Other including Epitaph and Escabrosa

Two samples of ground Horquilla sulfide ore were examined by detailed mineralogical modal analysis. The result of this analysis indicates that there is a large difference in copper mineralogy within the Horquilla rock type. Silver appears to be associated mainly with the copper sulfide minerals as is minor gold. Molybdenite, MoS2, is the only molybdenum mineral identified.

The copper oxide mineralization is principally chrysocolla, tenorite, malachite, and azurite. Oxide resources are distributed in three major rock units as follows:

  • Arkose
  • Porphyry – Quartz Monzonite (QMP) or Quartz Laterite (QLP)
  • Andesite

For the most part, core samples from exploration drilling were used for metallurgical testing. Split core samples were used for most of the comminution and some leach tests, while coarse rejects and split core were used for flotation testing. Whole core was used for some tests including the JK Drop-weight and impact crushing tests. Bulk surface samples were also taken for some of the column leach tests.

A fragmentation study was performed to predict the size distribution of ROM ore. The fragmentation study indicates that the ROM ore fed to the primary crusher will have a “Best Estimate” 80% passing size (P80) of about 30 inches, a size distribution readily handled by a size (60" x 110") crusher.

The comminution test program consisted of:

  • JK Drop-weight and Abrasion Test
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  • MinnovEX SAG Power Index Test (SPI)
  • MacPherson Autogenous Grindability Test
  • Bond Low-energy Impact (Crushing) Test
  • Bond Rod Mill Work Index Test
  • Bond Ball Mill Work Index Test
  • Bond Abrasion Test
  • Specific Gravity Determination

Grinding mill sizing parameters were provided to mill manufacturers for use in their mill sizing methods. The mill sizing parameters are shown in Table 13-1.

Table 13-1: Grinding Mill Sizing Parameters

Parameter Value
CWi 4.90
RWi 12.40
BWi 11.40
Tonnage 3,400 tph
SAG Mill Feed Size 150,000 µ
Transfer Size 3,000 µ
Ball Mill Product Size 105 µ

Flotation test work was performed during the years 1974-1975 and 2006-2008. The tests included bench-scale rougher-scavenger and cleaner tests, rougher variability tests, and rougher cleaner optimization tests. Based on the test results the flotation conditions were indicated to be as follows:

  • Primary grinding to P80=105µ
  • Rougher flotation pH= 9.7 to 10.8
  • AP-238 and AX-343 collectors
  • Regrind to P80= 74µ
  • One stage of cleaner flotation

The rougher flotation variability tests examined the effect of grind size, ore grade, ore mineralogy, and ore depth on metal recovery. The result of the variability tests indicated that there is not a strong correlation between head grade, copper mineralogy (as determined by logging), and mining level and copper recovery in the samples tested. Previous early-stage testing determined that the degree of sample oxidation was the most significant factor in the metallurgical response.

The result of the variability tests indicated that the grind size has an effect on both copper recovery and rougher concentrate grade. The mineralogical modal analyses indicate that the chalcopyrite liberates at a coarser size, between 150 and 75µ, than do the bornite and chalcocite. The moly begins to liberate from the gangue between 150 and 75µ, but remains locked to a significant degree with gangue to about 22µ.

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In the variability tests, only about 10% of the samples gave molybdenum recovery of 75% or higher, indicating that the variability test conditions were probably not optimum for moly recovery. Normally a molybdenum recovery of about 80% can be expected with a typical southern Arizona copper rougher concentrate. The result of sorting the variability test result for molybdenum recovery and ore elevation indicates no correlation between these variables.

During 2008, flotation tests were conducted at MSRDI on composite samples of five individual rock lithology samples and one composite sample representing the ore expected to be processed during the first three years of process plant operation. The test program was designed to examine the process of producing molybdenite concentrate. The bulk (copper-molybdenite) flotation concentrate from Horquilla ore produced a molybdenite concentrate grading 52.7% molybdenum with a 93% molybdenum recovery from bulk concentrate. The results of testing the other samples indicate lower molybdenite concentrate grades and with variable molybdenite recovery from the bulk concentrate with the procedure used. The results of the testing are presented in Table 13-2.

Table 13-2: Molybdenite Flotation

Molybdenite Flotation 

Sample ID
Concentrate Assay % Recovery %
Mo
Cu Mo Insol
         
Horquilla          0.44 52.7 1.8 93.0
Colina          0.70 26.5 16.9 96.5
Earp          0.50 42.8 6.5 93.0
Epitaph          0.30 39.3 17.5 55.7
Escabrosa            0.50 27.9 25.8 84.8
1-3 Yr Composite          0.06 41.6 13.5 96.5

In 2012, a metallurgical test program was designed to prepare composite samples representing four periods of mine production and test them by bench scale test procedures. The test procedures followed the treatment methods proposed for the process plant. The metallurgical test composite samples were prepared from half-core drill hole segments from six holes drilled in late 2011. The drill core segments were selected so that the composite samples would have the ore grade, lithology, and spatial characteristics of ore predicted to be produced during the mine operating periods of years 1 through 3, years 4 through 7, years 8 through 12, and years 13 through 21.

The composition of the composite samples by lithology is shown in Table 13-3.

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Table 13-3: Lithology of Composite Samples





Lithology
Composite Samples Representing
Mine Production Years
1
through
3
4
through
7
8
through
12
13
through
21
         
Epitaph - - 10% 16%
Colina - 11% 17% 25%
Earp 16% 28% 23% 16%
Horquilla 84% 61% 50% 43%

The result of closed circuit flotation tests on the year 1 through year 3 composite sample indicates a final copper recovery of 87.9% and a molybdenum recovery of 62% in a final bulk concentrate grading 41% copper, 1.02% molybdenum, and 502 ppm silver.

The result of closed circuit flotation tests on the year 4 through year 7 composite sample indicates a final copper recovery of 81.2% and a molybdenum recovery of 2.5% in a final bulk concentrate grading 44% copper and 0.047% molybdenum. (The anomalous value for the molybdenum recovery was checked by re-testing the same composite sample and the results were a somewhat improved rougher concentrate molybdenum recovery (40 to 60%) but not as good as previously tested composite samples. Previous results from testing samples containing the Colina ore type indicated that lower molybdenum recovery was to be expected. The cause is not specifically known at this time.)

The result of closed circuit flotation tests on the year 8 through year 12 composite sample indicates a final copper recovery of 92% and a molybdenum recovery of 84% in a final bulk concentrate grading 28% copper and 1.22% molybdenum. Additional analysis of the concentrate produced in the test work indicates that the concentrate contained low amounts of contaminates such as arsenic (<80 g/t) and mercury (0.8 g/t) and contained payable amounts of gold (1.91 g/t) and silver (294 g/t).

The result of closed circuit flotation tests on the year 13 through year 21 composite sample indicates a final copper recovery of 75.8% and a molybdenum recovery of 31.1% in a final bulk concentrate grading 36% copper, and 0.56% molybdenum. The core submitted for the year 13 to year 21 composite contained a higher percentage of oxidized material than will be mined, which resulted in lower metal recovery. A second composite sample was compiled with less oxidized material. Results of the second sample closed cycle test indicates copper recovery of 91.4% and molybdenum recovery of 66% in a final bulk concentrate containing 37% copper and 0.84% molybdenum.

An estimate of metal production in concentrate for the first 21 years of plant operation was prepared from the results of flotation test work performed by MSRDI in 2009 and MSRDI, G&T, and SGS in 2012. The 2012 work indicated the bulk concentrate production that could be expected by treating the expected ore composition for the operating years 1 through 21. The 2009 work indicated the separation efficiency that could be expected from treating bulk (copper-molybdenite) concentrate to produce a molybdenite concentrate and a final copper concentrate.

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Graphical analysis determined the bulk concentrate copper and molybdenite recovery that can be expected when the concentrate grade is fixed at 30% copper. It has been estimated that the recovery of molybdenite to a molybdenite concentrate separated from the bulk concentrate will be 95%. The molybdenite production was then calculated by applying the separation efficiency factor (95%) to the estimated annual production of molybdenite in the bulk concentrate as determined by locked cycle flotation testing of the composite samples.

Silver recovery was determined by submitting the flotation products from locked cycle flotation testing of the composite samples to fire assay procedures to determine the silver contents of each. The average silver recovery by annual periods was than estimated from the results of test and assays.

The estimates of annual metal recovery are presented in Table 13-4.

Table 13-4: Estimated Metal Recovery by Year of Production

Estimated Metal Recovery by Year of Production

Production Year
Recovery %
Cu Mo Ag
       
1 89.8 65.0 77.5
2 89.8 65.0 77.5
3 89.8 65.0 77.5
4 84.1 34.2 72.6
5 84.1 34.2 72.6
6 84.1 34.2 72.6
7 84.1 34.2 72.6
8 90.6 78.7 78.2
9 90.6 78.7 78.2
10 84.8 74.3 73.9
11 82.1 72.2 71.8
12 84.4 73.9 73.5
13 84.0 56.7 73.1
14 85.5 57.2 74.3
15 89.1 58.6 76.9
16 89.1 58.6 76.9
17 89.1 58.6 76.9
18 89.1 58.6 76.9
19 89.1 58.6 76.9
20 89.1 58.6 76.9
21 89.1 58.6 76.9

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Reagent consumption rates for the full scale plant operation have been estimated from test results. The estimated reagent consumption rates for sulfide ore processing are shown in Table 13-5.

Table 13-5: Estimated Reagent Consumption Rates

Estimated Reagent Consumption Rates

Item
Rate
lbs/ton ore
Copper Circuit  
         Aero Promoter 8944 0.04
         Collector, C-7 0.098
         Frother, Methyl Isobutyl Carbonal (MIBC) 0.026
         Lime 1.797
         Sodium Meta Silicate 0.14
         #2 Diesel Fuel 0.026


Item
Rate
lbs/ton
Copper-Moly Concentrate
Molybdenite Circuit  
         Sodium Hydrosulfide 934.4
         Sodium Meta Silicate 25.4
         #2 Diesel Fuel 15.2
         Methyl Isobutyl Carbonal (MIBC) 15.2
         Flomin D-910 88.9

Copper-moly and moly cleaner flotation tests indicate that the Rosemont sulfide ores should respond well to widely used and proven techniques. Reagent screening tests were performed that indicated recovery from the rock type composites could be improved by reagent selection.

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14                     MINERAL RESOURCE ESTIMATES

The updated Mineral Resource Estimate for the Rosemont deposit is prepared by Susan C. Bird of Mouse Mountain Technical Services (MMTS). This represents an update from the 2008 resource estimate (WLR Consulting, 2008) based on drilling and sample results up to March 2012, as well as updated geology.

The resource model is built using MineSight®, an industry standard in geologic modeling and mine planning software. The three dimensional block model has block dimensions of 50’x50’x50’ to reflect both the drill spacing and the bench height.

The block model limits, based on Imperial coordinates converted from UTM NAD 83 are: 1,710,000 to 1,722,000 East; 11,550,000 to 11,560,000 North; and 2,500 foot to 6,500 foot elevation. The volume modeled covers the extent of the main mineralized zone, as well as all pit limits tested.

The geology is separated into domains based on lithology and also into three metallurgical zones (oxide, mixed, and sulfide) based on interpretation of the assay data.

Statistical analyses (cumulative probability plots, histograms, contact plots and classic statistical values) of the assay and composite data are used to confirm the domain selection and to decide if capping is necessary within each zone and domain. Assay data is composited into 50’ intervals. The composites are used to create correlograms for Cu, Mo and Ag grades using the MSDA module of MineSight®, thus establishing rotation and search parameters for the block model interpolation. The composites used during interpolation are limited by both zone and domain. The resource is then classified as Measured, Indicated or Inferred based on variogram parameters and in accordance with the CIM Definition Standards (CIM, 2005).

Validation of the model is completed by comparison of the Ordinary Kriged (OK) values with both Inverse Distance Squared (ID2) and Nearest Neighbor (NN) interpolated block values, by the use of swath plots and grade tonnage curves. A visual inspection in section and plan throughout the deposit was performed to compare the modeled grades with the assay data. In addition, average grades for each data set were compared.

14.1                  DRILL HOLE DATABASE

Additional data since the 2008 resource estimate includes assays from 11 of the 12 drill holes recently completed by Augusta, as well as the additional sampling from five previously drilled exploration holes. One exploration drill hole, AR-2080, had not yet been assayed at the time of the model build. There are a total of 258 holes within the block model limits (of the 266 holes in the database), for a total number of intervals of 58,281with assayed Cu values.

All mineralized zones for every hole were drilled using diamond core, with less than five percent started by open-hole rotary techniques.

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14.1.1               Assay Statistics and Domain Definition

Assay statistics for each lithologic unit have been used to determine the domains for interpolation. Changes to the lithology since 2008 include the separation of all units above and below the Low Angle Fault (LAF). Lithologic boundaries have been re-interpreted where necessary, requiring minor changes due to additional drilling information. Coding of both lithology and domain to the model is done using 3D wireframe solids created in MineSight for each unit.

The assay data, by lithology, was analyzed to determine appropriate hard boundaries required to form domains for the interpolation. Assay means by domain, contact plots, as well as geologic knowledge on the location of major faults all contributed to the final domain assignments. Changes to the domain boundaries from the 2008 model include:

  1.

Combining lithologies Colina and Epitaph below the LAF

  2.

Combining lithologies Glance, Scherrer, Concha and Epitaph above the LAF

  3.

Combining Martin and Martin West

  4.

Separating the Horquilla and Earp formations

The resulting combinations of lithologies for each domain are summarized in Table 14-1.

Table 14-1: Domain Definition based on Lithology

Domain Lithologic
Units
Description
1
2
4
5
6
7
8
9
10
11
12
13
15
16
17
21
1
2,3
4
5
6
7,18
8
9
10
11,22,212,214
12,14
13
15
16
17
21
Overburden
Epitaph and Colina - below LAF
Earp
Horquilla,
Escabrosa
Martin, Martin West
QMP
Andesite
Arkose
Glance, Epitaph, Scherrer, Concha - above LAF
Scherrer, Concha - below LAF
Abrigo
Bolsa
Granodiorite
Epitaph North
Gravel

The resulting assay statistics for total copper (TCu) by domain and zone are summarized in the tables 14-2 through 14-5, and for Mo in tables 14-6 through 14-9. Examination of these tables indicates that within the sulfide and mixed zones, domains 2, 4, 5, 6, and 8 are the primary ore bearing domains. Within the oxides, only domains 8 through 10 contain lithology conducive to leaching. Of these, domains 8 and 9 are both significant leach oxide hosts.

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Table 14-2: Assay Statistics of TCu for Domains 8 through 10 – Oxides

  Domain
Parameter 8 9 10
Num Samples 915 837 8794
Num Missing Samples 0 0 0
TCu (%) Min 0 0 0
TCu (%) Max 11.900 3.890 3.500
TCu (%) - Weighted Mean 0.175 0.136 0.044
TCu (%) - Weighted SD 0.533 0.296 0.122
TCu (%) - Weighted Var 0.284 0.088 0.015
TCu (%) - Weighted CV 3.051 2.185 2.8

Table 14-3: Assay Statistics of TCu for Domains 1 through 9 – Sulfides

  Domain
Parameter 1 2 4 5 6 7 8 9
Num Samples 0 9352 5384 14178 1846 869 1030 1789
TCu (%) Min - 0 0 0 0 0 0 0
TCu (%) Max - 14.500 32.200 15.390 24.700 3.260 15.880 10.300
TCu (%) - Weighted Mean - 0.423 0.273 0.473 0.321 0.052 0.328 0.149
TCu (%) - Weighted SD - 0.636 0.356 0.7 1.535 0.205 0.85 0.428
TCu (%) - Weighted Var - 0.405 0.127 0.49 2.355 0.042 0.723 0.183
TCu (%) - Weighted CV - 1.504 1.305 1.48 4.785 3.967 2.592 2.864

Table 14-4: Assay Statistics of TCu for Domains 10 through 21 – Sulfides

  Domain
Parameter 10 11 12 13 15 16 17 21
Num Samples 3082 1388 325 1059 374 193 518 11
TCu (%) Min 0 0 0 0 0 0 0 0
TCu (%) Max 3.140 3.650 1.500 16.520 1.510 2.210 6.880 0.040
TCu (%) - Weighted Mean 0.074 0.072 0.055 0.285 0.09 0.083 0.139 0.001
TCu (%) - Weighted SD 0.207 0.272 0.158 0.741 0.16 0.162 0.356 0.006
TCu (%) - Weighted Var 0.043 0.074 0.025 0.55 0.026 0.026 0.127 0
TCu (%) - Weighted CV 2.8 3.783 2.857 2.603 1.772 1.942 2.567 5.102

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Table 14-5: Assay Statistics of TCu for Domains 1 through 21 – Mixed

  Domain
Parameter  4 5 8 10 11 12 13
Num Samples 12 1119 190 6 4 68 581
TCu (%) Min 0.15 0 0 0 0.23 0 0
TCu (%) Max 0.450 5.050 1.420 0.340 0.550 0.510 2.360
TCu (%) - Weighted Mean 0.32 0.635 0.215 0.118 0.335 0.191 0.444
TCu (%) - Weighted SD 0.099 0.475 0.222 0.138 0.129 0.08 0.349
TCu (%) - Weighted Var 0.01 0.226 0.049 0.019 0.017 0.006 0.122
TCu (%) - Weighted CV 0.309 0.748 1.031 1.165 0.385 0.418 0.786

Table 14-6: Assay Statistics of Mo for Domains 8 through 10 – Oxides

  Domain
Parameter 8        9 10
Num Samples 915 837 8794
TCu (%) Min 0 0 0
TCu (%) Max 0.144 0.195 0.105
Mo (%) - Weighted Mean 0.006 0.0022 0.0013
Mo (%) - Weighted SD 0.0086 0.0079 0.0039
Mo (%) - Weighted Var 0.0001 0.0001 0
Mo (%) - Weighted CV 1.4345 3.5698 3.1272

Table 14-7: Assay Statistics of Mo for Domains 1 through 9 – Sulfides

  Domain
Parameter 1 2 4 5 6 7 8 9
Num Samples 0 9352 5384 14178 1846 869 1030 1789
TCu (%) Min - 0 0 0 0 0 0 0
TCu (%) Max - 5.966 1.590 2.660 0.350 0.040 0.260 0.030
Mo (%) - Weighted Mean - 0.0127 0.0144 0.0157 0.0056 0.0028 0.0164 0.0019
Mo (%) - Weighted SD - 0.0668 0.0312 0.0463 0.0135 0.0047 0.0254 0.004
Mo (%) - Weighted Var - 0.0045 0.001 0.0021 0.0002 0 0.0006 0
Mo (%) - Weighted CV - 5.2528 2.1642 2.9568 2.4346 1.6665 1.5514 2.1133

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Table 14-8: Assay Statistics of Mo for Domains 10 through 21 – Sulfides

  Domain
Parameter 10 11 12 13 15 16 17 21
Num Samples 3082 1388 325 1059 374 193 518 11
TCu (%) Min 0 0 0 0 0 0 0 0
TCu (%) Max 0.132 0.302 0.126 0.430 0.155 0.030 0.170 0.006
Mo (%) - Weighted Mean 0.0016 0.0024 0.0044 0.0057 0.0021 0.0025 0.0059 0.0001
Mo (%) - Weighted SD 0.0045 0.0088 0.008 0.0201 0.0076 0.0042 0.0164 0.0006
Mo (%) - Weighted Var 0 0.0001 0.0001 0.0004 0.0001 0 0.0003 0
Mo (%) - Weighted CV 2.8479 3.639 1.8276 3.5339 3.6475 1.6447 2.7791 6.0051

Table 14-9: Assay Statistics of Mo for Domains 1 through 21 – Mixed

  Domain
Parameter 4 5 8 10 11 12 13
Num Samples 12 1119 190 6 4 68 581
TCu (%) Min 0.001 0 0 0 0 0 0
TCu (%) Max 0.011 0.560 0.050 0.114 0.001 0.012 0.111
Mo (%) - Weighted Mean 0.0047 0.0061 0.0048 0.0101 0.0005 0.0046 0.0033
Mo (%) - Weighted SD 0.003 0.0125 0.0053 0.0305 0.0005 0.0029 0.0123
Mo (%) - Weighted Var 0 0.0002 0 0.0009 0 0 0.0002
Mo (%) - Weighted CV 0.6508 2.0427 1.1059 3.0061 1 0.6358 3.7812

14.1.2               Compositing of Drill Hole Data

Compositing is done by 50’ bench composites in order to correspond to the planned bench height and elevations. The domains are coded to the composites by a majority code using the model block codes. The zones are coded using 3D surfaces corresponding to the bottom of each zone layer.

Composites are used to determine capping of metal values during the interpolations. Cumulative Probability Plots (CPP) are created for each metal and domain in the sulfide zone to determine that lognormal distribution applies, and to aid in selection of the capping required. In the mixed and oxide zone, each metal is plotted for all domains together due to the lack of data. Resulting capping values are given in Table 14-10.

Figures 14-1 and 14-2 are CPP plots for each of the three modeled metals (Cu, Mo and Ag) in the oxide and mixed zones respectively. Figures 14-3 through 14-5 are CPP plots of the domains requiring capping in the sulfide zone for Cu, Mo and Ag respectively.

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Table 14-10: Capping Value of Composites

Zone Domain Capping Threshold
TCu
(%)
Mo
(%)
Ag
(opt)
Sulfides 2
4
5
6
7
8
9
10
11
12
13
15
16
17
21
2
1.7
---
---
1.7
1.5
1.2
0.5
---
---
---
---
---
---
---
---
---
0.2
---
---
---
---
---
---
---
---
---
---
---
---
0.6
---
---
---
1
0.8
0.3
0.2
---
---
1.2
---
---
---
---
Mixed all --- --- 0.2
Oxides all 1.2 0.35 0.5

For grade values above the capping limits, the search distance for use of the value during interpolation is restricted to 50 ft. Beyond this distance, the capped value of the composite is used for interpolation.

Figure 14-1: CPP Plots of each Metal for All Domains – Oxide Zone

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Figure 14-2: CPP Plots of each Metal for All Domains – Mixed Zone

Figure 14-3: CPP Plots of TCu for Sulfide Domains requiring Capping

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Figure 14-4: CPP Plots of Mo for Sulfide Domains requiring Capping

Figure 14-5: CPP Plots of Ag for Sulfide Domains Requiring Capping

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14.2                  VARIOGRAPHY

In order to determine the directional properties of each domain, correlograms have been created for all three metals in each domain and zone. It was found that there is insufficient data to produce reliable correlograms by domain separately. Therefore, within the sulfide zone, the primary mineralized domains – domains 2, 4, and 5 are used to select the rotational and kriging parameters for all domains except the quartz monzonite porphyry (QMP) domain.

The lithologic units within the Rosemont deposit area strike generally approximately N-S and dip moderately to the east, as is evident in the geologic section of Figure 7-4. A spherical model is used to obtain the best fit in all cases. Variography adheres to the bedding, as is indicated in the summary of parameters listed in Table 14-11. The major axis of the spheroid plunges down-dip of the formations, with the minor axis plunging to the north at 15 degrees. The major and minor axes correlograms and corresponding spherical models are illustrated in Figures 14-6 and 14-7 for TCu.

The QMP intrusions did not indicate any directional preference, and therefore omni-directional correlograms are determined to be appropriate for this domain for each metal and zone.

Table 14-11: Variogram Parameters

Domains Zone Metal Rotation
(GSLIB-MS)
Axis Range 1
(ft)
Range
2 (ft)
Nugget Total
Sill
Sill1 Sill2
All except QMP All TCu
ROT
DIPN
DIPE
110
-55
15
Major
Minor
Vert
450
400
300


0.4
1
0.6
0
Mo
ROT
DIPN
DIPE
110
-55
10
Major
Minor
Vert
300
300
150


0.4
1
0.6
0
Ag
ROT
DIPN
DIPE
110
-55
10
Major
Minor
Vert
300
300
100*
1500
1000
300
0.4
1
0.42
0.18
QMP All TCu
Mo
Ag
Omni-Directional 400
300
300


0.4
0.2
0.3
1
1
1
0.6
0.8
0.7



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Figure 14-6: Major Axis Correlogram – TCu

Figure 14-7: Minor Axis Correlogram - TCu

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14.3                  BLOCK MODEL INTERPOLATION AND RESOURCE CLASSIFICATION

The block model built for the Rosemont deposit has the dimensions summarized in Table 14-12.

Table 14-12: Block Model Dimensions

Direction Minimum (ft) Maximum (ft) Block Dimension
(ft)
# of Blocks
Easting 1,710,000 1,722,000 50 240
Northing 11,550,000 11,560,000 50 200
Elevation 2,500 6,500 50 80

The block model is coded according to domain, based on wire-frame solids and zone surfaces using oxide and transition surfaces derived from the drill hole database information, including the acid soluble copper content.

The topographic surface is based on an aerial survey flown by Cooper Aerial Surveys Company of Tucson, Arizona in the summer of 2006. The vertical datum is based on the NAVD 88 standard. Cooper provided electronic files with elevation data on 10-foot contour intervals covering the project area. The percent of the block below topography is also calculated into the model blocks.

Specific gravity values are based on 392 measurements by Skyline Laboratories based on the differential of the weight in air and the weight in water. Table 14-13 summarizes the tonnage factors coded to the block model.

Table 14-13: Specific Gravity by Lithology

Lithology Rock Code Tonnage Factor (ft3/ton)
Overburden, unconsolidated 1 13.72
Epitaph Formation 2 12.11
Colina Limestone 3 11.69
Earp Formation 4 11.73
Horquilla Limestone 5 11.18
Escabrosa Limestone 6 11.56
Martin Formation 7 11.98
Quartz Monzonite Porphyry 8 12.31
Mesozoic Andesite 9 11.53
Willow Canyon Arkose 10 12.08
Glance Conglomerate/Ls 11 11.68
Scherrer Formation 12 12.00
Abrigo Formation 13 11.35
Concha Limestone 14 12.11
Bolsa Quartzite 15 11.91
Precambrian Granite 16 11.91
Epitaph North 17 12.11
Martin West 18 11.98
Undefined 19 12.00
Undefined 20 12.00
Tertiary Gravel 21 13.72

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The interpolation is completed using ordinary kriging (OK) in 2 passes with search parameters based on the variogram parameters. The first pass maximum search distances are equal to ½ the range of the variograms, and the second pass has a maximum search distance equal to the full range. Restrictions on the search distances and number of composites used in each pass are given in Table 14-14 and Table 14-15 below. The selection of a composite for interpolation is also restricted by both the domain and the zone codes, which are required to match the block model codes.

Table 14-14: Interpolation Search Parameters

Domains Zone Metal Rotation
(GSLIB-MS)
Axis Distance
(ft)
1st Pass
Distance
(ft)
2nd Pass




All Except
QMP






all




TCu
ROT
DIPN
DIPE
110
-55
15
Major
Minor
Vert
225
200
150
450
400
300

Mo
ROT
DIPN
DIPE
110
-55
10
Major
Minor
Vert
150
150
75
450
400
300

Ag
ROT
DIPN
DIPE
110
-55
10
Major
Minor
Vert
150
150
50
450
400
300

QMP

all
TCu
Mo
Ag
Omni-Directional 200
150
150
400
400
400

Table 14-15: Interpolation Composite Restrictions

Metal
Interpolation
Pass
Min. #
Comps
Max #
Comps
Max
Comps/DH
Max
Comps/Quad
TCu 1
2
3
1
8
8
2
1
2
4
Mo 1
2
3
1
12
12
2
1
2
4
Ag 1
2
3
1
8
6
2
1
2
4

14.3.1               Resource Classification

Classification of the resource into Measured, Indicated, and Inferred is based on the variogram parameters and restrictions on the number of composites and drill holes used in each pass of the interpolation. The resource is classified as Measured or Indicated according to the distances and composites numbers summarized in Table 14-16, with domain boundaries not honored for the purposes of classification. Inferred blocks are defined as all blocks with grades interpolated that do not meet the measured or indicated constraints.

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The definition of Indicated and Inferred used to classify the resource is in accordance with that of the CIM Definition Standards (CIM, 2005).

Table 14-16: Interpolation Composite Restrictions

Class Minimum # Comps Maximum Distance (ft)
Measured 3
1
100
50
Indicated 2 320
Inferred 1
450
(400 for QMP)

14.4                  BLOCK MODEL VALIDATION

Validation of the model is completed by comparison of mean grades, swath plots, grade-tonnage curve comparisons, and visual inspection in section and plan across the extent of the model.

14.4.1               Comparison of Mean Grades

The following Table compares the block model interpolated values for the kriged grades (OK) and Nearest Neighbor (NN) for blocks within the resource pit. For the oxide, only the domains with leach potential are compared (domains 8-10). For the Sulfide zone, all domains are used in the comparison.

The Nearest Neighbor interpolation is essentially the composite data, de-clustered to remove any bias in the drilling locations. Each metal and zone indicates good correlation between the OK grade and the de-clustered composite data.

Table 14-17: Comparison of OK grades with NN Grades within Resource Pit

    CUOK CUNN Diff MOOK MONN Diff AGOK AGNN Diff
Zone Parameter (%) (%) (%) (%) (%) (%) (opt) (opt) (%)

Oxide
Samples 30815 30815 na 30815 30815 na 30815 30815 na
Missing Samples 199 199 na 199 199 na 199 199 na
Weighted Mean 0.047 0.049 -4.1% 0.0013 0.0012 4.2% 0.0076 0.0071 7.0%

Sulfide
Samples 85789 85789 na 85789 85789 na 85789 85789 na
Missing Samples 614 614 na 614 614 na 614 614 na
Weighted Mean 0.351 0.338 3.8% 0.0134 0.013 3.1% 0.0971 0.091 6.7%

14.4.2               Grade-Tonnage Curves

Grade-tonnage curves are used to compare distribution data of the interpolated OK grades with the Nearest Neighbor distributions and the Nearest Neighbor Corrected (NNC). The NNC model is used in order to correct for the change of sample size from the composite to the block size of 50’x50’x50’. The Indirect Lognormal Correction that has been used is based on the variogram parameters, the block size, the mean grades, and the Coefficient of Variation of the NN grades.

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Figures 14-8 through 14-16 illustrate the grade-tonnage curves for each metal and zone. For the oxide zone, only domains 8 through 10 are plotted, as they are the only leachable veins. Each plot indicates that the OK grades are slightly lower than the corrected NN (NNC) grades throughout the grade distribution, and particularly at the higher end of the curve. This is to be expected, and accounts for the internal smoothing of the model during interpolation of the OK grades. Search parameters were chosen in an iterative procedure in order to create tonnage-grade curves that correspond to the necessary amount of dilution expected. Average copper equivalent values for sulfides/mixed material above a cutoff grade of 0.20 percent indicate a copper grade that is 2 percent lower for the OK interpolation compared to the NNC interpolation. Because of the modeling procedure, no additional dilution is required when reporting the resource and reserves.

Figure 14-8: Grade-Tonnage Curve for TCu-Sulfide Zone

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Figure 14-9: Grade-Tonnage Curve for TCu-Mixed Zone

Figure 14-10: Grade-Tonnage Curve for TCu-Oxide Zone

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Figure 14-11: Grade Tonnage Curve for Mo-Sulfide Zone

Figure 14-12: Grade-Tonnage Curve for Mo-Mixed Zone

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Figure 14-13: Grade-Tonnage Curve for Mo-Oxide Zone

Figure 14-14: Grade-Tonnage Curve for Ag-Sulfide Zone

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Figure 14-15: Grade-Tonnage Curve for Ag-Mixed Zone

Figure 14-16: Grade-Tonnage Curve for Ag-Oxide Zone

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14.4.3               Swath Plots

Swath plots through the block model are created in the N-S, E-W, and vertical directions for the three metals, in order to compare the ordinary kriged (OK) grades to those of the Nearest Neighbor (NN) model, which acts as a proxy for the de-clustered data. Swath plots of the inverse distance squared (ID) interpolation are also plotted as a further check. These are illustrated in Figures 14-17 through 14-20. Each “swath” is created as an average of the block grades for each direction shown, with steps of 100 feet along the easting and northing plots, and 50 feet along the vertical plots. The bar graph of the block tonnage provides an indication of the location of the majority of the data. The swath plots do not indicate any global bias in the OK grade values and show good correlation with the NN and ID grades throughout the main body of the data.

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Figure 14-17: Swath Plots for TCu-Sulfide Zone

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Figure 14-18: Swath Plots for Mo – Sulfide Zone

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Figure 14-19: Swath Plots for Ag – Sulfide Zone

14.4.4 Visual Validation

A series of E-W, N-S sections and plans (every 100’) have been used to inspect the ordinary kriged block model grades with the drill hole data. Figures 14-20 through 14-22 give examples of this comparison for the E-W section at 11,554,825N, for Cu, Mo and Ag grades respectively. The drill hole projection is 100’ from the section. Figures 14-23 through 14-25 are plans of the kriged Cu, Mo, and Ag grade, respectively, along with the composite Cu grades at the 4,000 foot bench elevations. The following Figures include the zone surfaces for the bottom of the oxide and mixed zones as well as the resource and ultimate pit outline.

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Plots throughout the model confirmed that the block model grades corresponded very well with the assayed grades.

Figure 14-20: E-W Section at 11,554,825N of OK Model and Assays – TCu Grades

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Figure 14-21: E-W Section at 11,554,825N of OK Model and Assays – Mo Grades

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Figure 14-22: E-W Section at 11,554,825N of OK Model and Assays – Ag Grades

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Figure 14-23: Plan at 4000’ of OK Model and Assays – TCu Grades

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Figure 14-24: Plan at 4000’ of OK Model and Assays – Mo Grades

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Figure 14-25: Plan at 4000’ of OK Model and Assays – Ag Grades

14.5                  IN SITU MINERAL RESOURCE ESTIMATE

The mineral resource estimation work was performed by Susan C. Bird, M.Sc., P.Eng. a Senior Associate at MMTS and an independent Qualified Person under the standards set forth by NI 43-101. The final resource model was from May 25, 2012 and was referred to as the “712 model”.

A Lerchs-Grossman (LG) pit shell having a 45 degree slope angle has been applied to the three dimensional block model to ensure reasonable prospects of economic extraction for the reported mineral resources. Metal prices used for the resource pit are $3.50/lb Cu, $15/lb Mo and $20/oz Ag. The resource pit optimization was based on mining costs of $0.777/ton of mineralized material and $0.882/ton of waste material. For sulfide/mixed material a processing cost of $4.20/ton of mineralized material and a general and administrative (G&A) cost of $0.70/ton of mineralized material, for a total of $4.90/ton, was used. For oxide material a processing cost of $3.03/ton of mineralized material was used. These costs are in line with those developed for use in the mineral reserves.

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For the reporting of the in-situ resource by equivalent copper (EqvCu) within the LG pit shell, the metallurgic recoveries, metal prices, and resulting net smelter prices (NSPs) used, are summarized in Table 14-18.

Table 14-18: Base Case Recoveries, Metal Prices and Resulting Net Smelter Prices


Metal
Metal
Price
Oxides Mixed Sulfide
NSP Recovery NSP Recovery NSP Recovery
Cu $2.50 /lb $2.425 /lb 65% $2.078 /lb 40% $2.078 /lb 86%
Mo $15 /lb 0 0 $13.095 /lb 30% $13.095 / lb 63%
Ag $20 /oz 0 0 $17.111 /oz 38% $17.111/oz 80%

The equivalent copper grades are calculated based on the above information, resulting in the following equations for each metallurgical zone:

Sulfide: EqvCu% = Cu% + (Mo% * 0.63 * 13.095) + (AgOPT * 0.80 * 17.111)  
          (0.86 * 2.078)   (0.86 * 2.078 * 20)  
                 
Mixed: EqvCu% = Cu% + (Mo% * 0.30 * 13.095) + (AgOPT * 0.38 * 17.111)  
          (0.40 * 2.078)   (0.40 * 2.078 * 20)  
                 
Oxide: EqvCu% = Cu%          

The in situ resource is classified as Measured, Indicated or Inferred corresponding to Canadian National Instrument 43-101 standards (CIM, 2005). The resource by equivalent copper grade for the Rosemont deposit is summarized in Tables 14-18 through 14-22, for Measure, Indicated, Measured+Indicated, and Inferred mineral resources respectively. The tables present a range of cutoffs, of which the base case equivalent copper values for each zone are highlighted in each table. These cutoffs are sufficient to cover the processing plus G&A costs for the sulfide and mixed material ($4.90/ton) and the processing costs of the oxide material ($3.03/ton), at the expected metallurgical recoveries.

The measured and indicated mineral resource presented here is inclusive of the mineral reserve presented in the Mineral Reserve section. Mineral resources that are not mineral reserves do not have demonstrated economic viability.

Due to the uncertainty that may be associated with Inferred mineral resources it cannot be assumed that all or any part of inferred mineral resources will be upgraded to an Indicated or Measured resource.

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Table 14-19: Measured Resource by Cu Equivalent Grade

Table 14-20: Indicated Resource by Cu Equivalent Grade

 
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Table 14-21: Measured + Indicated Resource by Cu Equivalent Grade

Table 14-22: Inferred Resource by Cu Equivalent Grade

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Augusta’s 2012 drilling campaign at the Rosemont deposit has increased both the quantity and confidence level of the estimated mineral resources, which presently totals about 919.3 million tons of measured and indicated, sulfide and mixed mineral resources grading 0.51% CuEqv, 0.41% Cu, 0.014% Mo, and 0.11 ounces per ton Ag, at a 0.15% CuEqv cutoff for sulfide and 0.30% CuEqv cutoff for a minor mixed component. An additional 138.6 million tons of inferred sulfide and mixed mineral resources are estimated at a grade of 0.49% CuEqv, 0.40% Cu, 0.012% Mo, and 0.10 ounces per ton Ag, at the same cutoffs. Sulfide and mixed material can be combined as metallurgical test work of the mixed material indicates that it can be processed with the sulfide material to produce a concentrate. Augusta’s recent drilling program and resource modeling was successful in converting significant tonnages of material previously classified as inferred into measured and indicated resource.

In addition, geologic and metallurgical studies conducted by Augusta have shown the potential for considering the oxide copper mineralization that overlies the sulfide deposit. Estimated measured and indicated oxide mineral resources total 63.4 million tons grading 0.17% Cu, at a 0.10% CuEqv cutoff (for oxide % CuEqv = % Cu). An additional inferred oxide mineral resource of 1.1 million tons grading 0.15% Cu is present, using the same cutoff. Oxide material could potentially be processed by heap leaching, to recover the copper.

14.6                  ADDITIONAL MINERAL RESOURCE POTENTIAL

The classification of currently inferred sulfide and oxide mineral resources can potentially be improved with further drilling. Additional mineral resources may be found in extensions to the north and down-dip of the Rosemont Deposit. Mineralization is also known to occur at Broadtop Butte, which could potentially be added as a satellite development. Further mineralization also occurs in the Copper World and Peach-Elgin deposits on the Rosemont Property. The mineralized areas at Broadtop Butte, Copper World and Peach-Elgin are characterized by related styles of mineralization and have formed through common geologic processes as the Rosemont Deposit. Historic drilling by Anaconda, Anamax, and ASARCO intercepted significant copper grades in what are commonly widely spaced holes. These areas warrant further exploration and have the potential to add to the mineral resource base.

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15

MINERAL RESERVE ESTIMATES

   
15.1

MINING OVERVIEW

The Rosemont Deposit is a large tonnage, skarn-hosted, porphyry-intruded, copper-molybdenum deposit located in close proximity to the surface and is amenable to open pit mining methods. The proposed pit operations will be conducted from 50-foot-high benches using large-scale equipment, including: 12.25 -inch-diameter rotary blasthole drills, 60-cu-yd class electric mining shovels, 25 and 36-cu-yd front-end loaders, 46–cu-yd hydraulic shovel, 260 ton off-highway haul trucks, 580- to 850-hp crawler dozers, 500-hp rubber-tired dozers, 270- to 500-hp motor graders and 30,000-gallon off-highway water trucks.

The mine has a 21-year life, with sulfide ore to be delivered to the processing plant at an initial rate of 75,000 tons per day (tpd). Provisions are included to increase tonnage to 90,000 tons of ore per day in Year 12 of operations. Mine operations are scheduled for 24 hours per day, 365 days per year.

A pre-production period of 1.75 years, or 22 months, will be required to ensure ore is readily available at mill start-up. During this pre-production period, approximately 99 million tons of waste will be stripped and 6 million tons of ore will be moved to the ore stockpile. After mill start-up, an average of 143,000 tpd of waste rock must be removed to maintain adequate ore supplies for continuous plant operations, bringing the total daily material production from the open pit to about 225,000 tons.

All mineral reserve estimates and mine plans are based on the deposit model described in Section 14. Consistent with industry standards for feasibility-grade analyses, mineral reserves are based on only ore-grade material classified as proven and probable; all inferred mineral resources are treated as waste. Imperial units of measurement are used throughout the mine plans. Tons refer to short tons (2000 pounds) and “ktons” refer to tons x 1000.

15.2                  GEOTECHNICAL RECOMMENDATIONS

Call & Nicholas, Inc. (CNI) was contracted by Augusta in 2012 to provide an update of their geotechnical recommendations for slope angles for the open pit development of the Rosemont Deposit. The current and previous work included geologic and geotechnical mapping, drilling, rock strength testing and slope stability analysis to determine pit slope design criteria that is consistent with industry norms for safety and cost effectiveness. CNI provided a report in February 2008 - Feasibility-Level Geotechnical Study For The Rosemont Deposit, and subsequently updated the report with a memorandum letter – Preliminary Findings from Slope Stability Review, June 19, 2009. CNI provided updated pit slope recommendations in 2012, based on their recent assessment. CNI also issued a memorandum letter– Slope Angle for Planned Open Pit Mine South Wall Tertiary Gravel Slope, July 20, 2012, updating the pit slope recommendation in the tertiary gravels on the south highwall. The following paragraphs in Section 15.2 are authored by CNI.

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The slope design parameters are a function of rock type (lithology), rock strength, wall height, faulting, and bedding and/or structure orientations. CNI subdivided the pit into 14 design sectors, which are illustrated in Figure 15-1.

Figure 15-1: Pit Slope Design Sectors and Maximum Slope Angles

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Table 15-1 below summarizes by design sector the slope angle recommendations used for all mining phase/pit plans presented in this feasibility study.

Table 15-1: Pit Slope Angle Recommendations

Design
Sector
Sector
Location
Slope Angles Bench Face
Angle
Bench
Height (ft)
Interramp Overall
1
2
3
4
5
6
7
8
9
Regional Fault – NW
Limestone – North
Limestone – N to NE
Limestone – South
Limestone – West
Limestone – ENE
Limestone – Central
Bolsa – NW
Bolsa - West
42°
43°
44°
45°
46°
47°
48°
49°
50°
-
-
-
-
-
-
-
-
-
58°
60°
66°
65°
66°
66°
68°
68°
68°
50
100
100
100
100
100
100
100
100
10
11
12
13
Tertiary Gravels
Willow Canyon Frm
Willow Canyon Frm
Willow Canyon Frm
-
-
-
-
35°
42°
33°
35°
-
-
-
-
50
50
50
50

Generally, slopes in the limestones/skarns and the Bolsa formation can be double-benched (i.e., a catch bench every 100 vertical feet) and are designed using interramp guidelines. An exception to this is Sector 1, which is in proximity to a strong regional fault and where single benching on 50-foot intervals is recommended.

All alluvium/overburden and arkose (Willow Canyon Formation) slopes should be single-benched and are limited to overall angles that are functions of wall height and groundwater levels. At the time of CNI’s analysis, the groundwater elevation was estimated to be around 4,400 feet in the immediate vicinity of the proposed open pit. Consequently, interramp slope angle versus slope height graphs were developed for pit walls above and below this elevation. Figure 15-2 presents the interramp slope graph for arkose. These interramp angles were applied mostly to the design of the internal mining phases as CNI’s recommendations in Table 15-1 already incorporated slope height allowances for the ultimate pit.

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Figure 15-2: Arkose Interramp Slope Angles

15.3                  PIT OPTIMIZATION

Processing plans call for the treatment of mostly sulfide ores at a milling rate of 75,000 tpd and increased later on for expansions and increased plant operating availabilities. The mill will produce two concentrates for shipment to off-site smelters or roasters: a copper concentrate that will include recoverable silver and a molybdenum disulfide concentrate.

Lerchs-Grossman analyses were conducted using the Rosemont Deposit model (described in Section 14) to determine the ultimate pit limits and best extraction sequence for open pit mine design. Only mineral resources classified as measured or indicated were considered as potential ore in the Lerchs-Grossman analyses; all inferred resources were treated as waste.

An economic subroutine was developed to compute a Net Smelter Return (NSR) value for each block in the deposit model. This computer algorithm incorporates block grades, expected smelting/refining contracts (i.e., payables and deductions), metallurgical recoveries and projected market prices for each metal (Cu, Mo and Ag) to yield a net revenue value expressed in terms of US Dollars per ton. The subroutine also applies to mining, ore processing and general/administration costs to calculate a net dollar value per block, which includes adjustments for surface topography. Concurrently, an equivalent copper grade is computed and stored in the block model.

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15.3.1               Metallurgical Recoveries

Metal recoveries were derived from metallurgical test work conducted by Mountain States Research and Development, Inc. (MSRDI), SGS (Vancouver division) and G&T Metallurgical Laboratories (G&T) in Kamloops, British Columbia. These tests included: grinding and flotation testwork. The metallurgical test work is described more fully in Section 13.

Based on preliminary results early in this test work, Table 15-2 presents the metallurgical recoveries used in the Lerchs-Grossman evaluations and subsequent mineral reserve estimation. Only the three primary metals – copper, molybdenum and silver – were modeled and used in the revenue calculations. No recovery of molybdenum and silver from oxide ore is projected.

Table 15-2: Metallurgical Recoveries Used in Lerchs-Grossman Evaluations

Metal Oxide Ore Sulfide Ore Mixed Sulfide Ore
Copper
Molybdenum
Silver
65 %
-
-
86 %
63 %
80 %
40%
30%
38%

15.3.2               Economic Parameters

Table 15-3 summarizes the economic parameters and offsite costs used in the base-case Lerchs-Grossman evaluations of the Rosemont Deposit.

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Table 15-3: Base-case Lerchs-Grossman Economic Parameters

Metal Prices:  
   Copper (Cu) $ 2.50 / lb Cu
   Molybdenum (Mo) $ 15.00 / lb Mo
   Silver (Ag) $ 20.00 / troy oz
Operating Costs (excl oxide leaching):  
   Base ore mining $ 0.777 / ton
   Base waste mining $ 0.882 / ton
   Incremental haulage (below pit rim at 5050 ft elevation) $ 0.028 / ton / bench
   Sulfide ore milling & flotation $ 4.20 / ton ore
   General/administration $ 0.70 / ton ore
Oxide Copper Ore Processing:  
   Cu oxide freight & refining $ 0.00 / lb Cu
   Acid consumption 28.6 lbs acid / ton ore
   Cost of acid $ 0.07 / lb acid
Oxide Ore Process Cost $ 3.03 / ton ore
Copper Concentrate Processing:  
   Cu grade in concentrate 30 %
   Cu realization 96.5 %
   Cu concentrate transportation $ 75.00 / dry ton
   Cu concentrate treatment $ 55.00 / dry ton
   Cu refining $ 0.055 / lb Cu
   Ag realization 90.0 %
   Ag refining $ 0.40 / troy oz Ag
Molybdenum Concentrate Processing:  
   Mo grade in concentrate 50 %
   Mo realization 90.0 %
   Mo concentrate transportation $ 0.00 / dry ton
   Mo treatment & refining $ 0.00 / lb Mo
NSR royalty 3 %

The base input mining costs are estimated from the results derived from the 2009 Feasibility Study (staff salaries in 2009 mining costs are excluded, and included with the G & A costs for this study). When applied along with the increment bench costs to the material contained within the base-case Lerchs-Grossman pit shell, the average mining cost is nearly $1.11 per ton of material. Mining costs near the pit bottom – below 3750 level, will exceed $1.68 per ton in 2012 US Dollars.

Consistent with current market conditions, no price participation charges are included in the concentrate processing costs. NSR values are computed using the parameters in Table 15-2 and Table 15-3, and are incorporated into the following formula for sulfide ore:

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NSR, $/ton = [(Net CuPrice * CuGrade * CuRec)
      + (Net MoPrice * MoGrade * MoRec)
      + (Net AgPrice * AgGrade * AgRec)]

Where: Net CuPrice = Cu price in $/lb net of offsite costs
  Net MoPrice = Mo price in $/lb net of offsite costs
  Net AgPrice = Ag price in $/oz net of offsite costs
  CuGrade = Interpolated block Cu grade expressed in %
  MoGrade = Interpolated block Mo grade expressed in %
  AgGrade = Interpolated block Ag grade expressed in oz/ton
  CuRec = Cu process recovery
  MoRec = Mo process recovery
  AgRec = Ag process recovery

For the Base Case metal prices:

  NSR(sulf), $/ton = ($2.078/lb * CuGrade/100 * 86% * 2000 lb/ton) + ($13.095/lb *
    MoGrade /100 * 63% * 2000 lb/ton) + ($17.111/oz * AgGrade * 80%)
     
  NSR(mixed), $/ton = ($2.078/lb * CuGrade/100 * 40% * 2000 lb/ton) + ($13.095/lb *
    MoGrade /100 * 30% * 2000 lb/ton) + ($17.111/oz * AgGrade * 38%)

Similarly, the NSR formula for oxide ore is:

  NSR(ox), $/ton = (NetCuPrice *CuGrade * (1-Royalty) * CuRec
     
    = ($2.50 * CuGrade/100 * (1-3%) * 65% * 2000 lb/ton

Bulk tonnage factors are read from the block model and combined with volume adjustments for surface topography effects, if any, to determine block tonnages. For each Lerchs-Grossman case, net profit values are calculated for each model block by subtracting on-site operating costs (mining, ore processing and G&A) from the NSR value, then multiplying the result by the block tonnage.

15.3.3               Slope Angles

Overall slope angles used on the Lerchs-Grossman evaluations were derived from the geotechnical recommendations made by CNI for pit slope designs. The overall slopes were adjusted to accommodate CNI’s recommended slope angles and the anticipated placement of internal haulage ramps along the pit walls in certain design sectors. CNI provide slopes angles for each model block, and a slope code was assigned to the block representing each of the pit slopes. The slope codes and pit slopes are then read as input to the Lerchs-Grossman analysis. The resulting overall slope angles are summarized in Table 15-4.

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Table 15-4: Overall Slope Angles Used in Lerchs-Grossman Analyses

Slope Code in Block
Model
Slope Angle
1
2
3
4
5
6
7
8
9
10
11
12
28°
33°
35°
42°
43°
44°
45°
46°
47°
48°
49°
50°

15.3.4               Lerchs-Grossman Analyses

All Lerchs-Grossman analyses were restricted to prevent the pit shells from crossing the topographic ridge immediately west of the deposit. This was done to minimize visual impacts from the Tucson metropolitan area.

The base-case Lerchs-Grossman pit shell is defined by the recoveries and economic parameters listed in Table 15-2 and Table 15-3, respectively. The metal prices of $2.50 /lb Cu, $15.00 /lb Mo and $20.00 /oz Ag are below a three-year trailing average. This pit shell contains about 755 million tons of measured and indicated sulfide mineral resources above an internal NSR cutoff of $4.90/ton and approximately 62 million tons of measured and indicated oxide mineral resources above a $3.03 /ton NSR cutoff. The resulting stripping ratio is about 1.8:1 (tons waste per ton of ore). However, this is not the pit shell selected for design. The current design for the tailings facilities has a limited capacity for approximately 680 million tons of ore feed, and a pit design around the base-case economic pit shell would produce tailings that will exceed the capacity of the storage facility. Therefore the selected pit shell has to contain measured and indicated sulfide mineral resources less than 680 million tons, and is picked from a set of simulations resulting from a sensitivity analysis.

Additional Lerchs-Grossman runs were made to evaluate sensitivities to metal prices and to mine operating costs. These sensitivities were generally conducted in 5% increments to +30% and –50% of the base case parameters. Table 15-5 and Table 15-6 present the results of the Lerchs-Grossman price and cost sensitivity analyses, respectively.

The pit shell that yields nearest to 680 million tons of measured and indicated mineral resource is one that is simulated with metal prices 25% less than the base case. The copper price used to generate this case is $1.88 /lb. It is selected as the basis for the ultimate pit design, and is approximately 9% smaller than the optimum economic pit.

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The selected Lerchs-Grossman pit shell compared to the base-case pit shell is shown in plan in Figure 15-3 and in cross section in Figure 15-4.

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Figure 15-3: Plan View Contours of Selected Lerchs-Grossman Pit Shell

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Figure 15-4: East West Section View of Selected Lerchs-Grossman Pit Shell

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Table 15-5: Lerchs-Grossman Results – Metal Price Sensitivities

      Sulfide Mineral Resources* Oxide Mineral Resources*      
  Prices Internal Cutoffs Above Internal NSR Cutoffs Above Internal NSR Cutoffs Waste Total Strip
  Cu Mo Ag       NSR TCu Mo Ag            
Sensitivity $/lb $/lb $/oz Sulfide Oxide Ktons $/t % % oz/t Ktons NSR $/t TCu % Ktons Ktons Ratio
                                 
+30% 3.25 19.50 26.00  4.90 3.03 785,495 18.55 0.42 0.014 0.12 67,391 5.40 0.17 1,628,490 2,481,376 1.91
+20% 3.00 18.00 24.00  4.90 3.03 779,526 18.62 0.42 0.014 0.12 67,329 5.40 0.17 1,601,579 2,448,434 1.89
+15% 2.88 17.25 23.00  4.90 3.03 776,396 18.66 0.42 0.014 0.12 67,318 5.40 0.17 1,586,457 2,430,171 1.88
+10% 2.75 16.50 22.00  4.90 3.03 768,862 18.74 0.42 0.014 0.12 67,308 5.40 0.17 1,549,866 2,386,036 1.85
+5% 2.63 15.75 21.00  4.90 3.03 762,080 18.82 0.43 0.014 0.12 67,298 5.40 0.17 1,525,459 2,354,837 1.84
Base 2.50 15.00 20.00  4.90 3.03 755,360 18.88 0.43 0.014 0.12 67,277 5.40 0.17 1,494,128 2,316,765 1.82
-5% 2.38 14.25 19.00  4.90 3.03 747,728 18.95 0.43 0.014 0.12 67,277 5.40 0.17 1,458,514 2,273,519 1.79
-10% 2.25 13.50 18.00  4.90 3.03 735,512 19.03 0.43 0.014 0.12 67,277 5.40 0.17 1,397,931 2,200,720 1.74
-15% 2.13 12.75 17.00  4.90 3.03 717,206 19.16 0.43 0.014 0.12 67,277 5.40 0.17 1,316,674 2,101,157 1.68
-20% 2.00 12.00 16.00  4.90 3.03 707,591 19.28 0.44 0.015 0.12 67,266 5.40 0.17 1,296,982 2,071,839 1.67
-25% 1.88 11.25 15.00  4.90 3.03 687,923 19.38 0.44 0.015 0.12 67,256 5.40 0.17 1,204,093 1,959,272 1.59
-30% 1.75 10.50 14.00  4.90 3.03 662,867 19.58 0.45 0.015 0.12 67,227 5.40 0.17 1,131,895 1,861,989 1.55
-35% 1.63 9.75 13.00  4.90 3.03 638,135 19.76 0.45 0.015 0.12 67,205 5.40 0.17 1,066,663 1,772,003 1.51
-40% 1.50 9.00 12.00  4.90 3.03 597,389 20.00 0.46 0.015 0.12 67,154 5.40 0.17 959,114 1,623,657 1.44
-45% 1.38 8.25 11.00  4.90 3.03 489,084 20.68 0.47 0.015 0.13 67,099 5.40 0.17 736,291 1,292,474 1.32
-50% 1.25 7.50 10.00  4.90 3.03 417,535 21.19 0.48 0.015 0.13 66,311 5.41 0.17 616,534 1,100,380 1.27

* Only measured and indicated mineral resources are reported above; all inferred mineral resources are treated as waste.

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Table 15-6: Lerchs-Grossman Results – Cost Sensitivities

    Internal   Oxide Mineral Resources*      
    NSR Sulfide Mineral Resources* Above Internal NSR      
  Prices Cutoffs Above Internal NSR Cutoffs Cutoffs Waste Total Strip
  Cu Mo Ag       NSR TCu Mo Ag     TCu      
Sensitivity $/lb $/lb $/oz Sulfide Oxide Ktons $/t % % oz/t Ktons NSR $/t % Ktons Ktons Ratio
                                 
+50% 2.50 15.00 20.00 4.90 3.03 725,441 19.07 0.43 0.014 0.12 67,277 5.40 0.17 1,336,228 2,128,946 1.69
+40% 2.50 15.00 20.00 4.90 3.03 735,814 19.01 0.43 0.014 0.12 67,277 5.40 0.17 1,390,396 2,193,487 1.73
+30% 2.50 15.00 20.00 4.90 3.03 737,969 19.00 0.43 0.014 0.12 67,277 5.40 0.17 1,399,143 2,204,389 1.74
+20% 2.50 15.00 20.00 4.90 3.03 746,815 18.93 0.43 0.014 0.12 67,277 5.40 0.17 1,443,712 2,257,804 1.77
+15% 2.50 15.00 20.00 4.90 3.03 750,421 18.91 0.43 0.014 0.12 67,277 5.40 0.17 1,462,073 2,279,771 1.79
+10% 2.50 15.00 20.00 4.90 3.03 750,617 18.91 0.43 0.014 0.12 67,277 5.40 0.17 1,462,245 2,280,139 1.79
+5% 2.50 15.00 20.00 4.90 3.03 752,313 18.90 0.43 0.014 0.12 67,277 5.40 0.17 1,476,334 2,295,924 1.80
Base 2.50 15.00 20.00 4.90 3.03 755,360 18.88 0.43 0.014 0.12 67,277 5.40 0.17 1,494,128 2,316,765 1.82
-5% 2.50 15.00 20.00 4.90 3.03 760,160 18.84 0.43 0.014 0.12 67,277 5.40 0.17 1,517,840 2,345,277 1.83
-10% 2.50 15.00 20.00 4.90 3.03 760,510 18.83 0.43 0.014 0.12 67,287 5.40 0.17 1,518,915 2,346,712 1.83
-15% 2.50 15.00 20.00 4.90 3.03 761,749 18.82 0.43 0.014 0.12 67,308 5.40 0.17 1,525,957 2,355,014 1.84
-20% 2.50 15.00 20.00 4.90 3.03 764,873 18.79 0.43 0.014 0.12 67,308 5.40 0.17 1,543,290 2,375,471 1.85
-25% 2.50 15.00 20.00 4.90 3.03 766,943 18.78 0.42 0.014 0.12 67,308 5.40 0.17 1,559,164 2,393,415 1.87

* Only measured and indicated mineral resources are reported above; all inferred mineral resources are treated as waste.

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The estimates presented in Tables 15-5 and 15-6 should not be confused with mineral reserves, which are based on open pit designs that incorporate access, operating, geotechnical and other criteria in addition to economic constraints. The Lerchs-Grossman results should not be relied upon, but do provide an indication of potential mineral reserves that must be validated by proper designs. Mineral resources that are not mineral reserves do not have demonstrated economic viability.

A 20% increase in metal prices boosts the ore-grade measured and indicated mineral resources by only 3%, while a 20% decrease in prices reduces them by about 6%. Figure 15-5 is a graphical presentation of the sensitivity to metal prices. Similarly, a 20% mine operating cost increase lowers ore-grade mineral resources by 1% and a 20% cost decrease adds to these resources also by about 1% in tonnage. If mine operating costs increase as much as 50%, there is only a 4% decrease in the contained resource. The graph in Figure 15-6 shows the sensitivity to mine operating costs.

In summary, the potential recoverable open pit resource is not very sensitive to metal prices until they drop below 40% less than the base case. It is even less sensitive to changes in mine operating costs. Upside pit expansion is impacted more as a result of the easterly dipping mineralized beds and the resulting rapidly increasing incremental stripping ratios.

Figure 15-5: Sensitivity Analysis on Metal Prices

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Figure 15-6: Sensitivity Analysis on Mine Operating Costs

15.4                  MINING PHASE DESIGNS

The ultimate Rosemont pit is designed for large-scale mining equipment (specifically, 60-cu-yd class electric shovels and 260-ton haulage trucks) and is derived from the selected Lerchs-Grossman pit shell described in the previous section. The design process included smoothing pit walls, eliminating or rounding significant noses and notches that may affect slope stability, and providing access to working faces by developing internal ramps.

15.4.1               Pit Design Parameters

The slope angles used for the design of the Rosemont ultimate pit and internal mining phases were presented earlier in Table 15-1 (see Section 15.2) . These slope angles allow for catch bench widths of 50-53 feet in the limestones/skarns and Bolsa Formation where the pit slopes are double-benched (i.e., vertical catch bench intervals of 100 feet). Slopes will be single-benched (i.e., on 50-foot intervals) in alluvium and arkose rock types, providing catch bench widths – toe to crest – of 25 to 48 feet. Interramp slopes and, hence, catch bench widths in alluvium and arkose vary according to the slope height and presence of groundwater.

The remaining parameters used in the designs of the ultimate pit and mining phases are presented in Table 15-7.

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Table 15-7: Pit Design Parameters

Bench height
Bench face angle
Catch bench interval – alluvium & arkose
Catch bench interval – all other rock types
50 ft
58-68°
50 ft
100 ft
Road width (including ditch & safety berm)
Nominal road gradient
125 ft
10 %
Minimum pushback width 300 ft

Mining phase, or pushback, widths are typically in excess of 300 feet, although operating widths are occasionally reduced to about 250 feet in limited areas. To maximize the ore recovery at the bottom of the ultimate pit, ramps are reduced to a single 70-foot lane (with berm and ditch) and maximum gradients are increased to 12%.

15.4.2               Mining Phases and Ultimate Pit

Seven mining phases define the extraction sequence for the Rosemont Deposit. The phase development strategy consists of the extracting the highest metal grades along with minimum strip ratios during the initial years to maximize the economic benefits of the ore-body, while enabling smooth transitions in waste stripping throughout the life of mine to ensure availability of ore feed to the mill.

The starter pit, Phase 1, is fit approximately to the Lerch-Grossman pit shell defined by a $1.09/lb Cu price (the 43% of base metal price sensitivity case). The set of pit shells that were used to approximate this phase as well as the subsequent ones is in Appendix D. This pit is located about 3,200 feet west of the primary crusher and ranges in elevation from 5,650 to 4,350 feet. The phase is approximately 2,600 feet wide east-west and 3,300 feet north-south. The upper benches will be dozed down until haul road access can be developed to the 5,550 foot elevation. Phase 1 will develop approximately 61 million tons of sulfide ore at a stripping ratio of 2.3:1 (tons waste per ton of total ore). An illustration of the Phase 1 pit is shown in Figure 15-7.

Phase 1 material will be accessed via a haul road that will be constructed from the pit exit eastward to the primary crusher. This road will also branch off towards the waste rock storage (WRS) areas. These roads will be used for the life of the project, and will also be extended to access the dry stack tailings areas.

The pit entrance is at the 5,150 foot elevation, and a ramp from that location enters the pit in a counter clockwise direction. The ramp switches back at the 4,950, and 4,650 foot elevations before reversing to a clockwise direction to the bottom of the pit. The benches below 4,550 foot elevation are access by a single lane with haul road.

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Figure 15-7: Plan View of Mining Pit Phase 1

Mining Phase 2 will expand the pit roughly 600 feet to the north, 400 feet to the east and 500 feet to the southeast. The eastern most limits of this pushback lie about 2,800 feet west of the primary crusher. Bench toe elevations will range from 5,750 to 4,300 feet. The phase is 2,900 feet wide east-west and 4,000 feet north-south. Phase 2 will supply over 27 million tons of sulfide ore. The average stripping ratio for this pushback is 3.1:1. An illustration of the Phase 2 pit is shown in Figure 15-8.

The pit entrance is at the 5,150 foot elevation, and a ramp from that location enters the pit in a clockwise direction. The ramp switches back at the 5,050, 4,750, and 4,550 foot-elevations before reversing to a counter clockwise direction to the bottom of the pit.

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Figure 15-8: Plan View of Mining Pit Phase 2

The open pit is further expanded 300 to 400 feet to the east with the development of Phase 3. The eastern most limits of this pushback lie about 2,500 feet west of the primary crusher. Benches will range between 5,750 and 4,150 feet toe elevations. The phase is 3,300 feet wide east-west and 4,200 feet north-south. Over 42 million tons of sulfide ore will be generated by Phase 3 at an average stripping ratio of 1.4:1. Phases 2, and 3 fit approximately to the Lerchs-Grossman pit shell defined by a $1.13/lb Cu price (the 45% of base case metal price sensitivity).

This expansion from the Phase 1 pit is split into 2 separate pushbacks both in the same general direction. For each phase expansion the ramp on the east side of the pit is re-developed. An illustration of the Phase 3 pit is shown in Figure 15-9.

The pit entrance is at the 5,150 foot elevation, and a ramp from that location enters the pit in a counter clockwise direction. The ramp switches back at the 5,050, 4,750, 4,550 and 4,400 foot elevations before reversing to a counter clockwise direction to the bottom of the pit.

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Figure 15-9: Plan View of Mining Pit Phase 3

Phase 4 will expand the open pit about 600 feet to the east and 400 feet to the north. The eastern most limits of this pushback lie about 1,700 feet west of the primary crusher. Phase 4 benches range between 5,450 and 3,950 feet. The phase is 3,800 feet wide east-west and 4,700 feet north-south. Phase 4 will produce nearly 43 million tons of sulfide ore at a stripping ratio of 2.4:1. Phase 4 is fit approximately to the Lerchs-Grossman pit shell defined by a $1.17/lb Cu price (the 47% of base case metal price value sensitivity). An illustration of the Phase 4 pit is shown in Figure 15-10.

The pit entrance is at the 5,100 foot elevation, and a ramp from that location enters the pit in a counter clockwise direction. The ramp switches back at the 4,950, 4,650, 4,450, 4,300 and 4,150 foot elevations before reversing to a clockwise direction to the bottom of the pit.

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Figure 15-10: Plan View of Mining Pit Phase 4

Phase 5 is fit approximately to the Lerchs-Grossman pit shell defined by a $1.28/lb Cu price (the 51% of base case metal price value sensitivity). Mining Phase 5 expands the pit approximately 300 feet to the north and 600 feet to the east. The eastern most limits of this pushback lie about 1,200 feet west of the primary crusher. Phase 5 benches range between 5,650 and 3,750 feet. The phase is 4,500 feet wide east-west and 4,800 feet north-south. Phase 5 will produce nearly 80 million tons of sulfide ore at a stripping ratio of 2.0:1. The ramp on the east side of the pit is redeveloped for this phase. An illustration of the Phase 5 pit is shown in Figure 15-11.

The pit entrance is at the 5,050 foot elevation, and a ramp from that location enters the pit in a counter clockwise direction. The ramp switches back at the 5,000, 4,800, 4,400, and 4,100 foot elevations before reversing to a counter clockwise direction to the bottom of the pit.

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Figure 15-11: Plan View of Mining Pit Phase 5

Phase 6 is fit approximately to the Lerchs-Grossman pit shell defined by a $1.46/lb Cu price (the 58% of base case metal price value sensitivity). Mining Phase 6 expands the pit by various distances in all directions. This pushback put the pit at the ultimate design limits along the north, west and south sides. The eastern most limits of this pushback lie less than 1,000 feet west of the primary crusher. The top of Phase 6 is 6,000 feet, near the top of the west ridge. The waste material at the upper elevations will be dozed down to a bench at approximately 5,600 feet elevation that will be wide enough for loading and hauling. A haul road will be constructed on the north side of the pit to access the 5,600 bench in this phase. The bottom bench of Phase 6 is 3,350 feet. The phase is 5,500 feet wide east-west and 6,400 feet north-south. Phase 6 will produce over 241 million tons of sulfide ore at a stripping ratio of 1.7:1. The ramp on the east side of the pit is re-developed for this phase. An illustration of the Phase 6 pit is shown in Figure 15-12.

The pit entrance is at the 5,050 foot elevation, and a ramp from that location enters the pit in a clockwise direction. The ramp switches back at the 4,750 and 4,250 foot elevations before reversing to a clockwise direction to the bottom of the pit.

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Figure 15-12: Plan View of Mining Pit Phase 6

The final pushback, Phase 7, extends the open pit from 300 to 600 feet along the east side to its ultimate limits and down to its maximum depth at the 2,900 foot elevation. The ultimate pit will be about 6,000 feet wide east-west and 6,500 feet wide north-south. The west wall will be over 3,000 feet high, while the east wall height will reach over 2,200 feet. Phase 7 is fit approximately to the Lerchs-Grossman pit shell defined by a $1.88/lb Cu price (the 75% of base case metal price value sensitivity). Phase 7 will generate nearly 172 million tons of sulfide ore at a stripping ratio of 1.7:1. An illustration of the Phase 7 pit, or final pit, is shown in Figure 15-13.

Total sulfide ore reserves in the final pit are estimated to be 667 million tons and 1.9 billion tons of waste material. Approximately 65 million tons of mineralized oxide material, indicated to be economic, are contained in this pit and included with the waste material instead of ore for the purpose of this study. Facilities to process this mineralized material is currently assessed and a portion of it may be included as ore in the future.

The ultimate pit is currently under-optimized due to the capacity limitations of the tailings storage facility. Potential expansion of this facility in the future will allow the pit to be designed to its optimum at the metal prices for $2.50 /lb Cu, $15.00 /lb Mo and $20.00 /oz Ag.

Throughout the mining phase and final pit designs, internal ramps are generally placed in arkose and alluvium pit walls in order to break up the interramp wall heights and limit the overall slope angles to the geotechnical recommendations. This allows the remaining walls to be steeper on an overall basis as fewer ramps were required in these design sectors. An additional benefit is to keep the main internal haulage ramps away from the high, western wall, which has the steepest inter-ramp slope angles. This allows additional sulfide ore to be developed at the base of the west wall.

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Figure 15-13: Plan View of Mining Pit Phase 7 (Ultimate Pit)

15.5                  MINERAL RESERVES

Rosemont mineral reserves have been estimated from only measured and indicated mineral resources; all inferred resources have been treated as waste. Inferred mineral resources have a great amount of uncertainty as to their existence and as to whether they can be mined legally or economically. It cannot be assumed that all or any part of inferred mineral resources will ever be upgraded to a higher category.

The mining phase and ultimate pit designs were applied to the 3D block model of the deposit described in Section 14 to estimate contained tonnages and grades. All reserve estimates are reported in Imperial units.

15.5.1               Ore Definition Parameters

The base-case price and operating cost estimates presented in Table 15-3 are used as the economic envelope to define ore in the mineral reserve estimates. These parameters are restated in Table 15-8 below. All prices and costs are in US dollars.

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Mineralized oxide materials that are indicated to be economic in the optimized pit analysis are not included in the pit ore reserves for this study. Designs for leaching facilities and recovery plans for the oxide materials are being undertaken, and they may be included as ore in future studies. All oxide materials are currently included with the waste materials.

Table 15-8: Ore Definition Parameters

Metal Prices:  
   Copper (Cu) $ 2.50 / lb Cu
   Molybdenum (Mo) $ 15.00 / lb Mo
   Silver (Ag) $ 20.00 / troy oz
Operating Costs (excl. oxide leaching):  
   Base ore mining $ 0.777 / ton
   Base waste mining $ 0.882 / ton
   Incremental haulage (below pit rim at 5050 ft elevation) $ 0.028 / ton / bench
   Sulfide ore milling & flotation $ 4.20 / ton ore
   General/administration $ 0.70 / ton ore
Copper Concentrate Processing:  
   Cu grade in concentrate 30 %
   Cu realization 96.5 %
   Cu concentrate transportation $ 75.00 / dry ton
   Cu concentrate treatment $ 55.00 / dry ton
   Cu refining $ 0.055 / lb Cu
   Ag realization 90.0 %
   Ag refining $ 0.40 / troy oz Ag
Molybdenum Concentrate Processing:  
   Mo grade in concentrate 50 %
   Mo realization 90.0 %
   Mo concentrate transportation $ 0.00 / dry ton
   Mo treatment & refining $ 0.00 / lb Mo
NSR royalty 3 %

15.5.2               Material Densities

Bulk material densities, which vary by rock type, were read from values stored in the block model. These assignments are described in more detail in Section 14. Generally, rock tonnage factors range between 11.18 ft3/ton to 13.72 ft3/ton and average about 11.85 ft3/ton for the rock contained within the ultimate pit.

15.5.3               Dilution

The Rosemont Deposit is a well-disseminated polymetallic deposit that has large ore zones above the anticipated internal cutoff grade. With the planned bulk mining method, external ore dilution along the ore - waste contact edges is generally assessed to determine whether the feed grade from the run of mine production is adequately represented by those predicted from the resource block model.

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The sample compositing and block grade interpolation process used to construct the deposit block model was determined to have incorporated sufficient dilution, and hence, no additional internal or external dilution factors are applied. This is previously explained in Section 14.4.2 that provides an analysis comparing the Cu grades from ordinary kriging to nearest neighbor composites.

The resource model block dimensions are 50 feet X 50 feet X 50 feet. The interpolated metal grade is averaged for the entire block. When this mine commences operations, ore feed will be delineated by implementing a blasthole sampling program. Blasthole spacings will be smaller, 25 feet to 35 feet, than the resource block dimensions, thereby provide better definition than from the resource block model. Therefore, there will likely be some selectivity within the dimensions of each resource model block allowing for separation of the ore from waste that is not evident on a whole block basis. The actual feed grade may be higher than model block value as a result. If an ore-waste contact exists within a model block, dilution has already been applied by averaging the metal grades on a whole block basis.

15.5.4               Mineral Reserve Estimates

The mineral reserve estimates presented in this report were prepared by Mr. Robert Fong, P.Eng., Principal Mining Engineer for Moose Mountain Technical Services. Mr. Fong meets the requirements of an independent qualified person under NI 43-101 standards. The mineral reserve estimates are effective as of July 17, 2012.

Proven mineral reserves for the Rosemont Deposit are summarized by mining phase in Table 15-9 and probable mineral reserves are presented in Table 15-10. Table 15-11 lists the combined proven and probable mineral reserve estimates and waste rock for the Rosemont Deposit.

As previous discussed, the pit designed for this study is under-optimized. It reflects an optimum pit at metal price of $1.88 /lb, Cu $11.07 /lb Mo, and $14.87 /oz Ag. Proven and probable sulfide mineral reserves within the designed final pit total nearly 667 million tons grading 0.44% Cu, 0.015% Mo and 0.12 oz Ag/ton. There are 1.24 billion tons of waste materials, resulting in a stripping ratio of 1.9:1 (tons waste per ton of ore). Total material in the pit is 1.9 billion tons. Contained metal in the sulfide (proven and probable) mineral reserves is estimated at 5.88 billion pounds of copper, 194 million pounds of molybdenum and 80 million ounces of silver. No mineralized oxide materials are in the ore reserves, they are included with the waste materials.

Nearly 46% of the sulfide mineral reserves in the Rosemont ultimate pit are classified as proven and the remainder (54%) is considered probable. The classifications are based on the exploration drilling in the Rosemont Deposit. All of the mineral reserve estimates reported above are contained in the mineral resource estimates presented in Section 14.

For possible metallurgical considerations, the combined proven and probable mineral reserves are broken out by principal rock types and/or geologic formations in Table 15-12.

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The Rosemont ultimate pit contains approximately 24 million tons of inferred sulfide mineral resources that are above the $4.90/ton NSR cutoff value for sulfides. These resources are included in the waste estimates presented in Table 15-11and Table 15-12. Inferred mineral resources are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves. Inferred mineral resources have a great amount of uncertainty as to their existence and as to whether they can be mined economically. It cannot be assumed that all or any part of inferred mineral resources will ever be upgraded.

All of the mineral reserve estimates presented in this report are dependent on market prices for the contained metals, metallurgical recoveries and ore processing, mining and general/administration cost estimates. Mineral reserve estimates in subsequent evaluations of the Rosemont Deposit may vary according to changes in these factors. There are presently no other known mining, metallurgical, infrastructure or other relevant factors that may materially affect the mineral reserve estimates.

It should be noted that there is some local environmental and political opposition to the development of the Rosemont open pit copper mining project. However, the right to mine and extract the minerals is provided for under federal law and Rosemont has valid mines claims. Rosemont will need to acquire all applicable permits from primarily federal and state agencies. These permits fully address all environmental media and provide for public and stake holder input. (See Section 20)

Rosemont mineral reserves are on mostly patented and some unpatented lands owned by Augusta Resource Corporation. Notwithstanding the existence of a 3% NSR mineral royalty and the existence of local environmental and political groups opposing the development of the project as noted above, the estimates of mineral reserves are not encumbered by any known legal, title, taxation, socio-economic, marketing, political, or other relevant issues.

Detailed listings of combined proven and probable mineral reserves by bench, by phase are presented in Appendix E.

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Table 15-9: Proven Mineral Reserves by Phase

  Sulfides >= 4.90 $/ton NSR Cutoff
Phase Ktons      NSR $/t TCu % Mo % Ag oz/t
           
1 36,335              22.31 0.50 0.016 0.14
2 16,334              17.27 0.40 0.011 0.09
3 23,075              19.99 0.43 0.018 0.14
4 22,947              22.12 0.50 0.013 0.15
5 46,471              22.18 0.51 0.013 0.13
6 107,605              19.32 0.45 0.014 0.13
7 55,308              19.54 0.43 0.016 0.11
           
Total 308,075              20.29 0.46 0.015 0.12

(NSR values are based on metal prices of $2.50/lb Cu, $15.00/lb Mo and $20.00/oz Ag.)

Table 15-10: Probable Mineral Reserves by Phase

  Sulfides >= 4.90 $/ton NSR Cutoff
Phase Ktons NSR $/t TCu % Mo % Ag oz/t
           
1 25,211              22.49 0.50 0.016 0.14
2 10,835              17.20 0.40 0.011 0.09
3 19,343              18.64 0.37 0.023 0.12
4 19,752              20.88 0.48 0.013 0.13
5 33,374              20.90 0.48 0.012 0.12
6 133,872              16.92 0.40 0.013 0.11
7 116,744              18.98 0.42 0.015 0.12
           
Total 359,131              18.67 0.42 0.014 0.12

(NSR values are based on metal prices of $2.50/lb Cu, $15.00/lb Mo and $20.00/oz Ag.)

Table 15-11: Combined Proven and Probable Mineral Reserves by Phase

  Sulfides >= 4.90 $/ton NSR Cutoff Waste Total Material Strip
Phase Ktons      NSR $/t TCu % Mo % Ag oz/t Ktons        Ktons Ratio
                 
1 61,546              22.38 0.50 0.016 0.14 142,729 204,275 2.32
2 27,169              17.24 0.40 0.011 0.09 84,526 111,695 3.11
3 42,418              19.37 0.40 0.020 0.13 59,553 101,971 1.40
4 42,699              21.54 0.49 0.013 0.14 100,709 143,408 2.36
5 79,845              21.64 0.50 0.013 0.13 156,603 236,448 1.96
6 241,477              17.99 0.42 0.014 0.12 411,973 653,450 1.71
7 172,052              19.16 0.42 0.015 0.11 287,362 459,414 1.67
                 
Total 667,206              19.42 0.44 0.015 0.12 1,243,455 1,910,661 1.86

(NSR values are based on metal prices of $2.50/lb Cu, $15.00/lb Mo and $20.00/oz Ag.)

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Table 15-12: Combined Proven and Probable Mineral Reserves by Rock Formation

Rock Type / Sulfides >= 4.90 $/ton NSR Cutoff  Waste  Total Strip
Formation Ktons NSR $/t TCu % Mo % Ag oz/t  Ktons  Ktons Ratio
   Oxide - QMP / QLP           24,280 24,280  
   Oxide - Andesite           95,887 95,887  
   Oxide - Arkose           521,426 521,426  
   Oxide - Other           0 0  
   Sulfide - QMP / QLP 17,612 14.96        0.31 0.018 0.06 0 17,612 0.0
   Sulfide - Abrigo 16,554 13.13        0.31 0.005 0.10 100,512 117,066 6.1
   Sulfide - Concha 2,159 8.71        0.17 0.006 0.13 31,204 33,363 14.5
   Sulfide - Epitaph 72,730 19.42        0.44 0.013 0.10 29,610 102,340 0.4
   Sulfide - Colina 98,542 25.77        0.59 0.020 0.12 12,194 110,736 0.1
   Sulfide - Earp 111,704 13.52        0.29 0.014 0.07 22,629 134,333 0.2
   Sulfide - Horquilla 281,334 21.41        0.49 0.016 0.15 68,251 349,585 0.2
   Sulfide - Escabrosa 22,325 25.07        0.60 0.007 0.17 21,767 44,092 1.0
   Sulfide - Other 44,246 9.29        0.23 0.003 0.05 152,095 196,341 3.4
   Overburden           7,053 7,053  
   Tertiary Gravels           156,547 156,547  
Total 667,206 19.42        0.44 0.015 0.12 1,243,455 1,910,661 1.9

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16                     MINING METHODS

Mining sequence plans are developed to depict mining progress at regular intervals and to serve as the basis for a mine production schedule. The sequence plans are developed from the phase designs described in Section 15.4 and target a sulfide (mill) ore production base rate of 75,000 tpd. This rate is reduced in Year 1 for mill ramp-up, and increased later on for expansions and increased plant operating availabilities.

16.1                  PRODUCTION SCHEDULING CRITERIA

The operating and scheduling criteria used to develop the mining sequence plans are summarized in Table 16-1 below.

Table 16-1: Mine Production Schedule Criteria

Annual Sulfide Ore Production Base Rate
Daily Sulfide Ore Production Base Rate
27,375,000 tons
75,000 tons
Operating Hours per Shift
Operating Shifts per Day
Operating Days per Week
Scheduled Operating Days per Year
12
2
7
365
Number of Mine Crews 4

Pit and mine maintenance operations will be scheduled around the clock. Allowances for down time and weather delays have been included in the mine equipment and manpower estimations presented in Sections 16.7 and 16.14.

The mill ramp-up schedule used for Year 1 production targets is presented in Table 16-2. Quarterly mill production in Year 2 will average about 6.8 million tons, equivalent to an annual rate of 27.4 million tons.

Table 16-2: Mill Ramp-Up Schedule (Year 1)


Month
% of Full
Production
Monthly
Ktons
Quarterly
Ktons
1
2
3
40
53
73
930
1,120
1,705

3,755
4
5
6
87
100
100
1,950
2,325
2,250

6,525
7
8
9
100
100
100
2,325
2,325
2,250

6,900
10
11
12
100
100
100
2,325
2,250
2,325

6,900
Total Year 1 88 24,080 24,080

The annual mill throughput for the life of mine is shown in Table 16-3. An expansion is planned beginning in Year 5 when the daily throughput will gradually increase from 75,000 tpd to 88,000 tpd by Year 7. In Year 12, an increase in plant operating availability will boost the daily throughput rate to 90,000 tpd.

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Table 16-3: Annual Mill Throughput

Year Average Throughput, tpd
1 65,973
2 75,000
3 75,000
4 75,000
5 78,000
6 84,000
7 88,000
8 88,000
9 88,000
10 88,000
11 88,000
12 90,000
13 90,000
14 90,000
15 90,000
16 90,000
17 90,000
18 90,000
19 90,000
20 90,000
21 90,000

16.2                  MILL FEED CUT-OFF GRADE STRATEGY

An elevated cut-off grade strategy is implemented to bring forward the higher grade ore from the pit to the early part of the ore production schedule. Delivering higher grade ore to the mill in the early years will improve the net present value economics of the project.

NSR values are calculated for each block in the resource model to represent the net Cu, Mo, and Ag metal values. The pit reserves are estimated on a cut-off with an NSR value of $4.90 /t. This is the minimum value of mineralized material that will cover the processing and G&A costs, and is therefore reserved for mill feed. Applying an elevated cut-off for mill feed in a given period will result in assigning the ore blocks with NSR values between $4.90 /t and the elevated cut-off NSR value to the ROM stockpile. The stockpiled ore will be reclaimed later in the life of mine. It is necessary for ore blocks destined to the stockpile to have an NSR value high enough to also cover handling costs. The additional cost is estimated to be $0.40 /t, and therefore ore blocks that are destined to the ROM stockpile must have a minimum NSR value of $5.30 /t. Otherwise this material will be sent to the waste storage areas instead.

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The elevated cut-off value to determine the head grade to the mill is limited at the upper end. Typically, more ore will be mined as well as additional waste stripping to extract sufficient mill feed as the mill cut-off value is elevated. The required mine equipment capacity to move the additional materials will limit the mill cut-off value. The maximum NSR cut-off value applied in this plan is $12.00 /t.

16.3                  OVERBURDEN STRIPPING REQUIREMENTS

Mineral reserve tabulations by bench, by phase and a mine production scheduling program are used to analyze long-term stripping requirements for the Rosemont project. Elevation and phase order dependencies and sinking rate controls are used in conjunction with mill ore production targets and an internal NSR cutoff of $4.90/ton to simulate open pit mining. The program, through successive iterations, allows the user to examine waste stripping rates over the life of the mine and their impact on ore exposure and mill head grades.

The stripping analysis determined a minimum preproduction stripping requirement of approximately 99 million tons of waste. Approximately 6 million tons of sulfide ore will also be mined and stockpiled during this period. The estimated Year 1 waste stripping total is 88 million tons, and 70 million tons for Year 2. The annual waste stripping from Year 3 through Year 12 will average about 76 million tons per year to maintain adequate ore exposure levels for uninterrupted ore supplies to the mill. Waste stripping rates will decline to an annual average of 42 million tons for the next 5 year period, and drop to an average of 3 million tons for the last 5 production years as the final mining phase reaches near the pit bottom.

Preproduction stripping will be conducted over a 21-month time period and will ramp up according to the delivery of mining equipment (particularly electric shovels) and the hiring and training of work crews. The long-term and peak mining rates suggest the use of at least three large (60-cy class) electric shovels, two large (36-cy and 25-cy) front-end loaders and a hydraulic shovel (46-cy). The preproduction stripping ramp-up is based on the delivery of the front-end loaders and hydraulic shovel 21 months prior to mill startup. Delivery of the first operating shovel is anticipated 16 months prior to mill startup, with successive deliveries on three-month intervals until the last shovel is placed into production ten months before startup.

Mining crews would typically be expanded every one to two months to allow time for hiring and training. Crew efficiencies would start off at reduced levels and increase with experience. Table 16-4 summarizes the mine’s preproduction stripping capacity ramp-up schedule if all equipment were utilized.

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Table 16-4: Mine Production Estimated Stripping Capacity Ramp-Up Schedule

Preproduction
Month
Mine
Capacity
Quarterly
Capacity
  Ktons Ktons
1
2
3
785
1,623
2,356

4,765
4
5
6
2,759
3,084
4,404

10,247
7
8
9
4,877
5,161
6,550

16,588
10
11
12
6,963
7,185
8,794

22,943
13
14
15
9,112
10,172
9,697

28,980
16
17
18
10,329
10,672
10,286

31,286
19
20
21
10,850
10,580
10,994

32,424
Total 147,233 147,233

16.4                  WASTE ROCK AND TAILINGS STORAGE

Overburden and other waste rock encountered in the course of mining will be placed into a waste rock storage (WRS) area located to the south and southeast of the planned open pit and into the dry stack tailings area, where dewatered mill tailings will be placed behind waste rock containment buttresses. The dry stack tailings area is north of the WRS area and east-northeast of the pit. The WRS and dry stack tailings facilities are fully contained within the Barrel drainage basin. The general mine site layout is shown in Figure 16-1.

The dry stack tailings facility is divided into two components, Phase 1 to the north and Phase 2 to the south. The dry stack tailings and WRS facilities were designed by AMEC/Tetra Tech, and are described in more detail in Section 18.5 of this report. MMTS provided estimates for waste rock quantities contained in the open pit and generated plans showing its development through the life of mine

16.4.1               Waste Rock Storage Design Criteria

The design criteria for the WRS area and associated haul roads are summarized in Table 16-5 below.

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Table 16-5: Waste Rock Storage Design Criteria

Angle of Repose 37°
Swell Factor on dumps
Average Tonnage Factor (with swell)
30 %
15.4 ft3/ton
Swell Factor on roads and buttresses
Average Tonnage Factor (with swell)
25 %
14.8 ft3/ton

Tetra Tech generated the estimates on waste rock quantities that are required for construction of the dry stack tailings buttresses and other structures requiring waste rock from the mine during the preproduction periods. Construction material will be supplied by stripping operations in the pit unless an alternate source is closer by. Any “surplus” waste rock will be directed to the starter buttresses and waste rock piles in the WRS area.

One of the objectives in the early years of operation (specifically, Years 1 to 5) is to construct a series of starter and perimeter buttresses (screen berms) around the eastern and southern perimeters of the WRS, and Phase 1 Dry Stack areas to provide a visual barrier to most of the mine’s operations from views along State Highway 83 and nearby private land owners located southeast of the project. These starter buttresses will also allow regrading and revegetation of the WRS side slopes at much earlier time periods than with traditional mine waste rock stockpile construction. To the extent possible, the long haulage profiles to the starter buttresses are balanced with shorter profiles to internal waste rock stockpiles.

The WRS buttresses and internal stockpiles are designed to facilitate subsequent regrading and reclamation. Side slopes in the WRS area will be regraded to a maximum of 3:1 (horizontal:vertical) slopes. Buttress construction will consist of haul trucks end-dumping waste rock in 100 foot lifts at a 37° angle of repose. Buttress crests will be set back to allow simple dozing of the crests down to meet the target regraded slope angles.

16.5                  MINE PLAN

Mining sequence plans are developed on a quarterly basis from preproduction through to the end of Year 2, and on an annual basis through Year 10. The preproduction period consists of seven quarters, or 21 months. Additional plans include mining progress through the end of Year 10, Year 15, Year 19, and Year 21.3 (end of mining). The mine plan drawings for these periods are show in Figure 16-1 to Figure 16-27. The mine production schedule is summarized in Table 16-6. Tables showing the benches mined by pit phase for each period are in Appendix F.

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Figure 16-1: Mine Plan Site Layout

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Figure 16-2: Mine Plan end of Period Pre-Production Q1

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Figure 16-3: Mine Plan End of Period Pre-Production Q2

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Figure 16-4: Mine Plan End of Period Pre-Production Q3

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Figure 16-5: Mine Plan End of Period Pre-Production Q4

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Figure 16-6: Mine Plan End of Period Pre-Production Q5

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  Figure 16-7: Mine Plan End of Period Pre-Production Q6

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Figure 16-8: Mine Plan End of Period Pre-Production Q7

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Figure 16-9: Mine Plan End of Period Year 01Q1

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Figure 16-10: Mine Plan End of Period Year 01Q2

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Figure 16-11: Mine Plan End of Period Year 01Q3

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Figure 16-12: Mine Plan End of Period Year 01Q4

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Figure 16-13: Mine Plan End of Period Year 02 Q1

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Figure 16-14: Mine Plan End of Period Year 02Q2

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Figure 16-15: Mine Plan End of Period Year 02Q3

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Figure 16-16: Mine Plan End of Period Year 02Q4

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Figure 16-17: Mine Plan End of Period Year 03

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Figure 16-18: Mine Plan End of Period Year 04

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Figure 16-19: Mine Plan End Period Year 05

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Figure 16-20: Mine Plan End of Period Year 06

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Figure 16-21: Mine Plan End of Period Year 07

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Figure 16-22: Mine Plan End of Period Year 08

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Figure 16-23: Mine Plan End of Period Year 09

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Figure 16-24: Mine Plan End of Period Year 10

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Figure 16-25: Mine Plan End of Period Year 15

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Figure 16-26: Mine Plan End of Period Year 19

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Figure 16-27: Mine Plan End of Period Year 21.3

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