MacArthur Copper Project - Quaterra Resources Inc

M3-PN 110127 

May 23, 2012 

MacArthur Copper Project 

NI 43-101 Technical Report 

Preliminary Economic Assessment 

Lyon County, Nevada, USA 

REVISION 0 

Prepared For: 

Qualified Persons: 

Myron R. Henderson, P.E. 

Rex C. Bryan, Ph.D. 

Herbert E. Welhener, MMSA-QPM 

Richard W. Jolk, P.E., Ph.D. 

Mark A. Willow, M.Sc., C.E.M.#1832 

M3 Engineering & Technology Corporation ● 2051 West Sunset Road, Tucson, AZ 85704 ● 520.293.1488 

&

MACARTHUR COPPER PROJECT 

FORM 43-101F1 PRELIMINARY ECONOMIC ASSESSMENT 

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. 

• Mr. Myron R. Henderson. P.E.; Project Manager with M3Engineering & Technology 

Corporation; principal author of this technical report and responsible for Sections 1 

through 3, Sections 17 through 19, Sections 21 and 22, and Sections 24 through 26. 

• Dr. Rex C. Bryan, Ph.D.; Senior Geostatistician with Tetra Tech MM, Inc.; responsible 

for Sections 4 through 12, Section 14, and Section 23. 

• Mr. Herbert E. Welhener, MMSA-QPM; Vice President of Independent Mining 

Consultants, Inc.; responsible for Section 16 – Mining Methods. 

• Dr. Richard W. Jolk, P.E., Ph.D., Principal Mine Engineer, Metallurgical Engineer, and 

Certified Minerals Appraiser with Tetra Tech MM, Inc.; responsible for Section 13 - 

Mineral Processing and Metallurgical Testing. 

• Mr. Mark A. Willow, M.Sc., C.E.M., Practice Leader with SRK Consulting (U.S.), Inc.; 

responsible for Section 20 – Environmental Studies, Permitting and Social or Community 

Impact. 

This Technical Report is current as of May 23, 2012 

(Signed) “Myron R. Henderson” June 26, 2012 

Myron R. Henderson, P.E. Date 

(Signed) “Rex C. Bryan” June 26, 2012 

Rex C. Bryan, Ph.D. Date 

(Signed) “Herbert E. Welhener” June 26, 2012 

Herbert E. Welhener, MMSA-QPM Date 

(Signed) “Richard W. Jolk” June 25, 2012 

Richard W. Jolk, P.E., Ph.D. Date 

(Signed) “Mark A. Willow” June 25, 2012 

Mark A. Willow, M.Sc., C.E.M. #1832 Date 

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TABLE OF CONTENTS 

SECTION PAGE 



DATE AND SIGNATURES PAGE ................................................................................................... i 

TABLE OF CONTENTS.................................................................................................................... ii 

LIST OF FIGURES AND ILLUSTRATIONS ............................................................................... ix 

LIST OF TABLES ............................................................................................................................ xii 

LIST OF APPENDICES .................................................................................................................. xv 

1 SUMMARY ............................................................................................................................. 1 

1.1 PROPERTY DESCRIPTION AND OWNERSHIP ............................................................... 1 

1.2 HISTORY ....................................................................................................................... 2 

1.3 GEOLOGY AND MINERALIZATION .............................................................................. 2 

1.3.1 Geophysics ............................................................................................... 3 

1.4 EXPLORATION STATUS ................................................................................................ 4 

1.4.1 Exploration Drilling Program ............................................................... 4 

1.5 RESOURCE ESTIMATE .................................................................................................. 5 

1.5.1 Block Model Definition .......................................................................... 5 

1.5.2 Assay Database ....................................................................................... 6 

1.5.3 Compositing ............................................................................................ 6 

1.5.4 Geostatistical Analysis and Variography ............................................. 7 

1.5.5 Kriging and Resource Classification .................................................... 7 

1.5.6 Estimated Resources .............................................................................. 9 

1.6 METALLURGY ............................................................................................................ 11 

1.7 ECONOMIC ASSESSMENT ........................................................................................... 12 

1.8 CONCLUSIONS AND RECOMMENDATIONS................................................................. 13 

2 INTRODUCTION ................................................................................................................ 14 

2.1 GENERAL .................................................................................................................... 14 

2.2 PURPOSE OF REPORT ................................................................................................. 14 

2.3 SOURCES OF INFORMATION ...................................................................................... 14 

2.4 CONSULTANTS AND QUALIFIED PERSONS ................................................................ 15 

2.5 DEFINITION OF TERMS USED IN THIS REPORT ........................................................ 16 

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3 RELIANCE ON OTHER EXPERTS ................................................................................. 20 

4 PROPERTY DESCRIPTION AND LOCATION ............................................................ 21 

4.1 LOCATION .................................................................................................................. 21 

4.2 PROPERTY OWNERSHIP ............................................................................................. 21 

4.3 MINERAL TENURE AND TITLE .................................................................................. 21 

4.4 RELEVANT INFORMATION ......................................................................................... 22 

5 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE 

AND PHYSIOGRAPHY ...................................................................................................... 26 

5.1 ACCESSIBILITY ........................................................................................................... 26 

5.2 CLIMATE .................................................................................................................... 26 

5.3 LOCAL RESOURCES AND INFRASTRUCTURE ............................................................ 26 

6 HISTORY .............................................................................................................................. 28 

6.1 PROPERTY HISTORY .................................................................................................. 28 

6.2 HISTORICAL RESOURCES .......................................................................................... 30 

6.3 HISTORIC MINING ..................................................................................................... 30 

6.4 HISTORIC METALLURGICAL TESTWORK AND MINERAL PROCESSING .................. 30 

7 GEOLOGICAL SETTING AND MINERALIZATION ................................................. 31 

7.1 REGIONAL GEOLOGY ................................................................................................ 31 

7.2 LOCAL GEOLOGY ...................................................................................................... 33 

7.3 PROPERTY GEOLOGY ................................................................................................ 33 

7.3.1 Alteration ............................................................................................... 36 

7.4 MINERALIZATION ...................................................................................................... 38 

8 DEPOSIT TYPES ................................................................................................................. 40 

8.1 OXIDE ZONE EXPLORATION ..................................................................................... 43 

8.2 CHALCOCITE/OXIDE ZONE EXPLORATION .............................................................. 43 

8.3 PRIMARY SULFIDE ZONE EXPLORATION ................................................................. 43 

9 EXPLORATION ................................................................................................................... 45 

9.1 GEOPHYSICS ............................................................................................................... 45 

9.1.1 IP/Resistivity Surveys ........................................................................... 45 

9.1.2 Airborne Magnetic Surveys................................................................. 60 

10 DRILLING ............................................................................................................................ 63 

10.1 HISTORICAL ............................................................................................................... 63 

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10.2 EXPLORATION & DRILLING HISTORY ...................................................................... 63 

10.3 HISTORIC MINING ..................................................................................................... 68 

10.4 CURRENT DRILLING .................................................................................................. 68 

10.5 SURVEYING DRILL HOLE COLLARS .......................................................................... 69 

10.6 DOWNHOLE SURVEYS ................................................................................................ 71 

10.7 CURRENT DRILLING METHODS AND DETAILS ......................................................... 71 

10.8 REVERSE CIRCULATION DRILLING SAMPLING METHOD ....................................... 72 

10.9 CORE DRILLING SAMPLING METHOD ...................................................................... 73 

10.10 DRILLING, SAMPLING, AND RECOVERY FACTORS .................................................. 73 

10.11 SAMPLE QUALITY ...................................................................................................... 73 

11 SAMPLE PREPARATION, ANALYSES AND SECURITY ......................................... 75 

11.1 RC SAMPLE PREPARATION AND SECURITY ............................................................. 75 

11.2 CORE SAMPLE PREPARATION AND SECURITY ......................................................... 75 

11.3 SAMPLE ANALYSIS ..................................................................................................... 76 

11.4 LEACH ASSAY ANALYSIS ........................................................................................... 77 

11.5 QUALITY CONTROL ................................................................................................... 79 

11.6 REVIEW OF ADEQUACY OF SAMPLE PREPARATION, ANALYSES, AND 

SECURITY ................................................................................................................... 80 

12 DATA VERIFICATION ...................................................................................................... 82 

12.1 HISTORIC DATA CHECK ............................................................................................ 82 

12.2 CURRENT DATA CHECK ............................................................................................ 82 

12.2.1 Adequacy of Data ................................................................................. 84 

13 MINERAL PROCESSING AND METALLURGICAL TESTING ............................... 85 

13.1 OXIDE ORE COPPER EXTRACTION ........................................................................... 86 

13.2 OXIDE ORE ACID CONSUMPTION ............................................................................. 87 

13.3 TRANSITION ORE EXTRACTION AND ACID CONSUMPTION ..................................... 87 

13.4 LEACH CYCLE TIME .................................................................................................. 88 

13.5 LEACH SOLUTION APPLICATION RATE .................................................................... 88 

13.6 PAD HEIGHT ............................................................................................................... 89 

13.7 PLS FLOW RATE AND PLS GRADE .......................................................................... 89 

13.8 PARTICLE SIZE TO HEAP LEACH .............................................................................. 89 

13.9 HEAP LEACH DESIGN CRITERIA ............................................................................... 89 

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14 MINERAL RESOURCE ESTIMATES ............................................................................. 92 

14.1 INTRODUCTION .......................................................................................................... 92 

14.2 MACARTHUR RESOURCE ESTIMATION .................................................................... 93 

14.3 MACARTHUR BLOCK MODEL ................................................................................... 95 

14.4 ASSAY DATA ............................................................................................................... 98 

14.5 COMPOSITE DATA .................................................................................................... 104 

14.6 GEOSTATISTICAL ANALYSIS AND VARIOGRAPHY.................................................. 109 

14.7 KRIGING ................................................................................................................... 112 

14.8 KRIGING ERROR AND RESOURCE CLASSIFICATION .............................................. 118 

14.9 VALIDATION OF BLOCK MODEL: VISUAL AND STATISTICAL CHECKS ................ 122 

14.10 MINERAL RESOURCE ESTIMATE ............................................................................ 127 

15 MINERAL RESERVE ESTIMATES .............................................................................. 132 

16 MINING METHODS ......................................................................................................... 133 

16.1 GEOTECHNICAL PARAMETERS ............................................................................... 133 

16.2 DILUTION MODELING AND FACTORS ..................................................................... 133 

16.3 OPEN PIT MINING .................................................................................................... 133 

16.4 MINING SCHEDULE .................................................................................................. 143 

16.5 WASTE DUMPS ......................................................................................................... 147 

16.6 MINING EQUIPMENT ................................................................................................ 154 

16.7 MINE LABOR ............................................................................................................ 155 

16.8 MINE CAPITAL COSTS ............................................................................................. 156 

16.9 MINE OPERATING COSTS ........................................................................................ 156 

17 RECOVERY METHODS .................................................................................................. 158 

17.1 OVERVIEW OF PLANNED FACILITIES ...................................................................... 158 

17.2 HEAP LEACH PAD .................................................................................................... 158 

17.3 SOLVENT EXTRACTION ........................................................................................... 159 

17.4 ELECTROWINNING ................................................................................................... 160 

17.5 SULFURIC ACID PLANT ............................................................................................ 161 

17.6 POWER PLANT .......................................................................................................... 162 

17.7 ANCILLARY FACILITIES ........................................................................................... 162 

18 PROJECT INFRASTRUCTURE ..................................................................................... 164 

18.1 SITE LOCATION ........................................................................................................ 164 

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18.2 PROCESS BUILDINGS ................................................................................................ 164 

18.3 ANCILLARY BUILDINGS ........................................................................................... 164 

18.3.1 Administration Building .................................................................... 165 

18.3.2 Warehouse / Plant Maintenance Building ....................................... 165 

18.3.3 Analytical Laboratory ........................................................................ 165 

18.3.4 Mine Truck Shop ................................................................................ 165 

18.3.5 Change House ..................................................................................... 165 

18.3.6 Main Gatehouse .................................................................................. 165 

18.3.7 Fuel Storage and Dispensing ............................................................. 166 

18.4 ACCESS ROADS......................................................................................................... 166 

18.5 RAILROAD FACILITIES............................................................................................. 166 

18.6 POWER SUPPLY & DISTRIBUTION ........................................................................... 166 

18.7 WATER SUPPLY & DISTRIBUTION .......................................................................... 166 

18.8 WASTE MANAGEMENT ............................................................................................ 167 

18.9 SURFACE WATER CONTROL ................................................................................... 167 

18.10 TRANSPORTATION & SHIPPING............................................................................... 167 

18.11 COMMUNICATIONS .................................................................................................. 168 

19 MARKET STUDIES AND CONTRACTS ...................................................................... 170 

20 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR 

COMMUNITY IMPACT .................................................................................................. 172 

20.1 ENVIRONMENTAL LIABILITIES ............................................................................... 172 

20.2 PERMITS ................................................................................................................... 172 

20.2.1 Federal Permitting .............................................................................. 174 

20.2.2 State Permitting .................................................................................. 176 

20.2.3 Local Permitting ................................................................................. 178 

20.3 ENVIRONMENTAL STUDIES ...................................................................................... 178 

20.4 WASTE AND TAILINGS DISPOSAL ............................................................................ 179 

20.5 PROJECT PERMITTING REQUIREMENTS ................................................................. 179 

20.6 SOCIAL OR COMMUNITY RELATED REQUIREMENTS ............................................ 179 

20.7 MINE CLOSURE REQUIREMENTS ............................................................................ 180 

21 CAPITAL AND OPERATING COSTS ........................................................................... 182 

21.1 CAPITAL COST ......................................................................................................... 182 

21.1.1 Mine Capital Cost ............................................................................... 182 

21.1.2 SX/EW Capital Cost ........................................................................... 182 

21.1.3 Sulfuric Acid Plant Capital Cost....................................................... 185 

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21.1.4 Exclusions ............................................................................................ 188 

21.2 RECLAMATION COST ESTIMATE ............................................................................ 188 

21.3 OPERATING COST .................................................................................................... 189 

21.3.1 Mine Operating Cost .......................................................................... 190 

21.3.2 SX/EW Operating Cost ...................................................................... 190 

21.3.3 Sulfuric Acid Plant Operating Cost .................................................. 191 

21.3.4 General and Administrative Costs ................................................... 193 

22 ECONOMIC ANALYSIS .................................................................................................. 195 

22.1 INTRODUCTION ........................................................................................................ 195 

22.2 MINE PRODUCTION STATISTICS ............................................................................. 195 

22.3 HEAP LEACH PAD AND SX/EW PRODUCTION STATISTICS ................................... 195 

22.3.1 Cathode Shipping ............................................................................... 195 

22.4 CAPITAL EXPENDITURE ........................................................................................... 196 

22.4.1 Initial Capital ...................................................................................... 196 

22.4.2 Sustaining Capital .............................................................................. 196 

22.4.3 Working Capital ................................................................................. 196 

22.4.4 Salvage Value ...................................................................................... 196 

22.5 REVENUE .................................................................................................................. 197 

22.6 OPERATING COST .................................................................................................... 197 

22.7 TOTAL CASH COST .................................................................................................. 197 

22.7.1 Royalty ................................................................................................. 197 

22.7.2 Reclamation and Closure ................................................................... 198 

22.8 DEPRECIATION AND DEPLETION ............................................................................. 198 

22.9 TAXATION ................................................................................................................. 198 

22.9.1 Income Tax and Mineral Tax ............................................................ 198 

22.10 PROJECT FINANCING ............................................................................................... 198 

22.11 NET INCOME AFTER TAX ........................................................................................ 199 

22.12 NPV AND IRR .......................................................................................................... 199 

22.13 SENSITIVITIES........................................................................................................... 199 

23 ADJACENT PROPERTIES .............................................................................................. 203 

23.1 SINGATSE PEAK SERVICES PROPERTIES ................................................................ 203 

23.2 OTHER PROPERTIES ................................................................................................ 204 

24 OTHER RELEVANT DATA AND INFORMATION ................................................... 207 

24.1 RE-PROCESSING OF YERINGTON RESIDUALS ........................................................ 207 

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24.1.1 Introduction ........................................................................................ 207 

24.1.2 Residual Copper Resources ............................................................... 207 

24.1.3 Mining Methods .................................................................................. 209 

24.1.4 Capital Cost Summary ....................................................................... 210 

24.1.5 Operating Costs .................................................................................. 211 

24.1.6 Economic Analysis .............................................................................. 213 

25 INTERPRETATION AND CONCLUSIONS ................................................................. 215 

25.1 RESOURCES .............................................................................................................. 215 

25.2 MINING METHODS ................................................................................................... 215 

25.3 METALLURGY .......................................................................................................... 216 

25.3.1 Run-of-Mine Heap Leaching ............................................................. 216 

25.3.2 Spatial Variability of In-Situ Size Distribution ............................... 216 

25.3.3 Chemical Degradation of the Ore during Leaching ....................... 216 

25.3.4 Permeability and Agglomeration ...................................................... 217 

25.3.5 Spatial Variability of Copper Extraction and Acid 

Consumption ....................................................................................... 217 

25.3.6 Relationship of Total Iron Mineralization to Acid 

Consumption ....................................................................................... 217 

25.4 ECONOMIC ASSESSMENT ......................................................................................... 217 

25.5 RISKS ........................................................................................................................ 218 

26 RECOMMENDATIONS ................................................................................................... 219 

26.1 METALLURGY TEST PROGRAM .............................................................................. 219 

26.1.1 Stage I- Sample Preparation ............................................................. 219 

26.1.2 Stage II- Acid Bottle Roll and Acid Characterization Testing ...... 219 

26.1.3 Stage III- Small Column Leach Tests .............................................. 220 

26.1.4 Stage IV- Large Column Leach Tests .............................................. 220 

26.1.5 Stage V- Study Preparation and Recommendations for a 

Final Feasibility ................................................................................... 220 

26.2 BUDGET AND SCHEDULE .......................................................................................... 220 

27 REFERENCES.................................................................................................................... 222 

APPENDIX A: CERTIFICATE OF QUALIFIED PERSON (“QP”) AND CONSENT 

OF AUTHOR ...................................................................................................................... 224 

APPENDIX B: PROPERTY LISTING ........................................................................................ 240 

APPENDIX C: EXPLORATION HISTORY OF THE MACARTHUR OXIDE 

COPPER PROPERTY ....................................................................................................... 255 

APPENDIX D: EXPLORATION DRILL HOLES WITH INTERCEPTS ............................. 260 

APPENDIX E: RESOURCE MODEL DRILL HOLE LISTING ............................................ 300 

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LIST OF FIGURES AND ILLUSTRATIONS 

FIGURE DESCRIPTION PAGE 

Figure 1-1: Quaterra Exploration Drilling by Year .........................................................................4 

Figure 4-1: General Location Map ................................................................................................23 

Figure 4-2: Regional Layout May..................................................................................................24 

Figure 4-3: MacArthur Property May ............................................................................................25 

Figure 6-1: Major Physiographic Features ....................................................................................29 

Figure 7-1: Regional Geology .......................................................................................................32 

Figure 7-2: Generalized Alteration Types .....................................................................................38 

Figure 8-1: Datamine© View of Resource Block Model Looking West ......................................40 

Figure 8-2: East-West Section 14,691,000N (Looking North) ......................................................41 

Figure 8-3: North- South Section 2,438,324 (Looking West) .......................................................42 

Figure 9-1: IPR line locations over the central MacArthur Project area. ......................................47 

Figure 9-2: Line 4300 (304300E) IP pseudo-section and inverted phase/depth model .................48 

Figure 9-3: Line 4300 Resistivity pseudo-section and inverted resistivity/depth model...............49 

Figure 9-4: Line 4900 IP pseudo-section and inverted phase/depth model ...................................50 


Figure 9-6: Line 7500 IP pseudo-section and inverted phase/depth model ...................................52 


Figure 9-8: QM-164 down hole electrode to remote electrode transmitter pair ............................54 

Figure 9-9: QM-177 down hole electrode to remote electrode transmitter pair ............................55 

Figure 9-10: Line location of the 1960’s Kennecott lines (in black) and the 2009 replacement line 

(in white). ...............................................................................................................57 

Figure 9-11: Historic and 2009 IP data on a modeled magnetic susceptibility depth slice ...........58 

Figure 9-12: Inversion model and pseudo-sections for line 6075 recorded in 2009. ....................59 

Figure 9-13: Location of the 2012 detailed helicopter magnetic survey .......................................61 

Figure 10-1: Location of Historic Drill holes ................................................................................65 

Figure 10-2: Drill hole Location Map ............................................................................................70 

Figure 10-3: Quaterra Exploration Drilling by Year .....................................................................72 

Figure 10-4: Letter from Mr. Henry Koehler .................................................................................74 

Figure 11-1: MacArthur Check Assay Results ..............................................................................80 

Figure 11-2: Reviewing Established Protocol for Data Entry .......................................................80 

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Figure 11-3: Manually Creating Geologic Sections from the Drill Data .......................................81 

Figure 12-1: Twin Hole Charted Results .......................................................................................84 

Figure 13-1: Comparison of Grade versus Copper Recovery Oxide Leach Ore ...........................86 

Figure 14-1: Drill Location and Search Zones for the MacArthur 2011 Model ............................95 

Figure 14-2: Side-by-Side Histograms – TCu% Assay SE-PIT area and NW-OUT area ...........104 

Figure 14-3: Side-by-Side Histograms – TCu% Composites SE-PIT area & NW area ..............109 

Figure 14-4: 0.12% Indicator Variograms (Omni Direction) For NW-Out and SE-Pit Areas ....110 

Figure 14-5: Selected Cu% Correlograms For SE-Pit And NW-Out Areas ................................111 

Figure 14-6: Side-by-Side Histograms M&I vs INF for (a) SE and (b) NW-Out .......................117 

Figure 14-7: Probability plot of kriging error ..............................................................................119 

Figure 14-8: Jackknife Method of Model Validation ..................................................................120 

Figure 14-9: Jackknife validation of kriging model (SE Area, MinZones 10 and 11) ................121 

Figure 14-10: Side-by-Side Samples, Composites and Blocks ...................................................122 

Figure 14-11: East West Cross Section Looking North (Cu blocks) ...........................................123 

Figure 14-12: East-West Cross Section Looking North (Resource Class) ..................................124 

Figure 14-13: North-South Cross Section Looking West (Cu Blocks) .......................................125 

Figure 14-14: North South Cross Section Looking West (Resource Class) ................................126 

Figure 16-1: Final Pits .................................................................................................................136 

Figure 16-2: Mining Phase 1 in MacArthur Pit ...........................................................................137 


Figure 16-4: Mining Phase 3 in North Pit Area ...........................................................................139 

Figure 16-5: Mining Phase 4 in North Pit Area ...........................................................................140 

Figure 16-6: Mining Phase 5 (Gallagher Pit) ...............................................................................141 


Figure 16-8: Final Pit and Dumps (including pit backfill) ..........................................................148 

Figure 16-9: End of Year 1 ..........................................................................................................149 

Figure 16-10: End of Year 3 ........................................................................................................150 



Figure 16-13: End of Year 10 ......................................................................................................153 

Figure 17-1: Overall Process Flowsheet ......................................................................................163 

Figure 18-1: MacArthur Heap Leach and Process Facilities .......................................................169 

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Figure 19-1: Historic Copper Price ..............................................................................................170 

Figure 22-1: MacArthur Project NPV Sensitivities .....................................................................200 

Figure 23-1: Adjacent Properties .................................................................................................206 

Figure 24-1: Yerington Mine Residuals ......................................................................................214 

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LIST OF TABLES 

TABLE DESCRIPTION PAGE 

Table 1-1: Exploration Drilling History ..........................................................................................2 

Table 1-2: MacArthur Model Parameters ........................................................................................5 

Table 1-3: MinZone Codes and Density ..........................................................................................6 

Table 1-4: Kriging and Search Parameters ......................................................................................8 

Table 1-5: Measured + Indicated Copper Resources .....................................................................10 

Table 1-6: Inferred Copper Resources ...........................................................................................11 

Table 10-1: Historic Exploration Drilling......................................................................................63 

Table 10-2: U.S. Bureau of Mines 1947-1950 Drilling Highlights ...............................................64 

Table 10-3: Anaconda Company 1955-1957 Drilling Highlights .................................................66 

Table 10-4: Pangea Exploration 1987-1991 Drilling Highlights ...................................................67 

Table 11-1: Sequential Copper Leach Assay Results ....................................................................78 

Table 11-2: Ferric Sulfate Leach (QLT) Assay Results ................................................................79 

Table 11-3: MacArthur 2011 QA/QC Program Results ................................................................79 

Table 12-1: List of Twin Holes Drilled By Quaterra .....................................................................83 

Table 13-1: MacArthur Historical Test Work ...............................................................................91 

Table 14-1: MacArthur Model Parameters ....................................................................................95 

Table 14-2: MinZone Codes and Density ......................................................................................96 

Table 14-3: MinZone Interval Data Count and Drill hole Assay Statistics ...................................97 

Table 14-4: Statistics of Cu Assay Data (All Areas) .....................................................................99 

Table 14-5: SE-Pit Area Cu Assay Statistics ...............................................................................102 

Table 14-6: NW Area TCu Assay Statistics ................................................................................103 

Table 14-7: MinZone Composite Count (All Areas) ...................................................................105 

Table 14-8: All Cu Assay Statistics for Quaterra Composites ....................................................106 

Table 14-9: SE Area Cu Assay Statistics for Quaterra Composites ............................................107 

Table 14-10: NW Area Cu Assay Statistics for Quaterra Composites ........................................108 

Table 14-11: Variogram and Search Parameters .........................................................................112 

Table 14-12: MinZone Block Count (All Areas) .........................................................................113 

Table 14-13: SE-Pit and NW Areas Cu Block Statistics .............................................................114 

Table 14-14: SE Area Cu Block Statistics ...................................................................................115 

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Table 14-15: NE Area Cu Block Statistics ..................................................................................116 

Table 14-16: Measured Copper Resources ..................................................................................128 

Table 14-17: Indicated Copper Resources ...................................................................................129 

Table 14-18: Measured + Indicated Copper Resources ...............................................................130 

Table 14-19: Inferred Copper Resources .....................................................................................131 

Table 16-1: Pit Definition Inputs .................................................................................................134 

Table 16-2: Floating Cone Geometries Used for Pit Designs......................................................135 

Table 16-3: Phase Tonnage and Grade Available for Mine Production Schedule ......................135 

Table 16-4: Production Schedule .................................................................................................143 

Table 16-5: Ore Production Schedule by Mining Phase ..............................................................144 

Table 16-6: Ore Production Schedule by Mining Phase and Resource Classification ................145 

Table 16-7: Waste Tonnage by Source and Destination ..............................................................147 

Table 16-8: Mine Equipment .......................................................................................................155 

Table 16-9: Mine Capital Estimate ..............................................................................................156 

Table 16-10: Mine Operating Costs .............................................................................................157 

Table 18-1: Products & Consumables .........................................................................................168 

Table 20-1: Summary of Major Permits for Future Mining ........................................................173 

Table 20-2: Future Baseline Studies ............................................................................................176 

Table 21-1: SX/EW Capital Cost .................................................................................................182 

Table 21-2: SX/EW Sustaining Capital .......................................................................................183 

Table 21-3: Sulfuric Acid Plant Capital Cost ..............................................................................186 

Table 21-4: Reclamation Cost Estimate ......................................................................................189 

Table 21-5: MacArthur SX/EW and Mine Operating Cost .........................................................190 

Table 21-6: SX/EW Operating Cost ............................................................................................190 

Table 21-7: Reagent Cost.............................................................................................................191 

Table 21-8: Sulfuric Acid Plant Operating Cost ..........................................................................192 

Table 21-9: General & Administrative Cost Summary ...............................................................193 

Table 21-10: General & Administrative Labor Cost Summary ...................................................194 

Table 22-1: Life of Mine Ore, Waste Quantities, and Ore Grade ................................................195 

Table 22-2: Initial Capital ............................................................................................................196 

Table 22-3: Life of Mine Operating Cost ....................................................................................197 

Table 22-4: Economic Indicators .................................................................................................199 

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Table 22-5: Sensitivity Analysis ..................................................................................................199 

Table 22-6: Discounted Cash Flow Model ..................................................................................201 

Table 23-1: Singatse Peak Services, LLC – Yerington Mine Resources, Feb. 2012 ..................203 

Table 23-2: Yerington Mine Residual Copper Resources, SRK March, 2012 (Non NI43-101 

Compliant) ...........................................................................................................204 

Table 23-3: Adjacent Property Resource Estimates ....................................................................205 

Table 24-1: Yerington Residual Oxide Copper Resources, SRK March 2012 ............................209 

Table 24-2: Combined Yerington Oxide Residuals / MacArthur Mine Capital & Sustaining 

Costs .....................................................................................................................211 

Table 24-3: Combined Yerington Oxide Residuals / MacArthur Mine Operating Costs............212 

Table 24-4: Combined Yerington Oxide Residuals / MacArthur Mine Economic Indicators ....213 

Table 26-1: Budget for MacArthur Follow on Test Work ...........................................................221 

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APPENDIX DESCRIPTION 

LIST OF APPENDICES 

A Certificate of Qualified Person (“QP”) and Consent of Author 

B Property Listing 

C Exploration History of the MacArthur Oxide Copper Property 

D Exploration Drill Holes with Intercepts 

E Resource Model Drill Hole Listing 

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1 SUMMARY 

Quaterra Alaska, Inc. (Quaterra), a wholly owned subsidiary of Quaterra Resources, Inc. 

commissioned M3 Engineering and Technology Corporation (M3) to prepare a Canadian 

National Instrument 43-101 (NI 43-101) compliant Preliminary Economic Assessment (PEA) for 

the MacArthur Copper Project in Lyon County, Nevada. Tetra Tech Inc. (Tt) and Independent 

Mining Consultants, Inc. (IMC) prepared several sections of the PEA. This report includes an 

update of the January 2011 MacArthur technical report and reflects changes to the resource 

estimate as a result of the 2011 exploration drilling and continued geologic investigations. 

The Qualified Person for Sections 4 through 12, Section 14, and Section 23 of this report is Mr. 

Rex Bryan, PhD, Senior Geostatistician for Tetra Tech, Golden Colorado. The Qualified Person 

for Section 13 of this report is Mr. Richard W. Jolk, P.E., PhD, Principal Minerals Engineer for 

Tetra Tech, Golden Colorado. The Qualified Person for Section 16 of this report is Mr. Herb 

Welhener, Principal Mining Engineer for Independent Mining Consultants, Inc., Tucson, 

Arizona. 

The MacArthur Copper Property is located near the geographic center of Lyon County, Nevada, 

USA along the northeastern flank of the Singatse Range approximately seven miles northwest of 

the town of Yerington, Nevada. The property is accessible from Yerington by approximately five 

miles of paved roads and two miles of maintained gravel road. Topographic coverage is on US 

Geological Survey “Mason Butte” and “Lincoln Flat” 7.5’ topographic quadrangles. The nearest 

major city is Reno, Nevada approximately 75 miles to the northwest. 

The Preliminary Economic Assessment within this Technical Report is based upon the oxide / 

chalcocite portion of the updated resource. This oxide / chalcocite portion includes a measured 

and indicated resource of 159.1 million tons averaging 0.21% Cu (percent total copper or TCu) 

containing 676 million pounds of copper at a 0.12% Cu cutoff and an inferred resource of 243 

million tons averaging 0.20% Cu at a 0.12% Cu cutoff containing 980 million pounds of copper. 

It should also be noted that integration of some 120 million tons of resource piles (non-compliant 

NI 43-101 “Residuals”) from Quaterra Resources 2011 acquisition of the historic, neighboring 

Yerington copper mine, could provide a significant positive impact on the economics of the 

MacArthur Project. Residuals consist of oxide-copper bearing sub-grade material representing 

stripped material from the Yerington mine, vat leach tailings representing oxide tailings from 

copper oxide vat leaching, and partially leached tailings and ore previously mined by Arimetco. 

The residuals are currently being characterized to elevate to a NI 43-101 status. 

1.1 PROPERTY DESCRIPTION AND OWNERSHIP 



the town of Yerington, Nevada. The property is accessible from Yerington by approximately five 

miles of paved roads and two miles of Lyon County maintained gravel road. Topographic 

coverage is on US Geological Survey “Mason Butte” and “Lincoln Flat” 7.5’ topographic 

quadrangles. The nearest major city is Reno, Nevada approximately 75 miles to the northwest. 

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The property consists of 470 unpatented lode claims totaling approximately 9700 acres on lands 

administered by the US Department of Interior - Bureau of Land Management (BLM). All 

required annual payments to the BLM and Lyon County have been paid in a timely manner and 

the claims are current. 

1.2 HISTORY 

Over the history of the project, previous operators have contributed more than 300 holes to the 

current drill hole database. Table 1-1 summarizes the exploration history of the MacArthur area. 

Of the historic holes, 280 of those holes drilled by the Anaconda Company (Anaconda) during 

1972-73 have been deemed acceptable under NI 43-101 standards and have been used during the 

resource estimation. 

Operator 

Table 1-1: Exploration Drilling History 

MACARTHUR PROJECT 

February 2009 

Drill Program 

Date Range 

Number of 

Holes Drilled 

Feet Drilled 

U.S. Bureau of Mines 1947-50 8 3,414 

Anaconda Company 1955-57 14 3,690 

Bear Creek Mining Company 1963-?? ~14 Unknown 

Superior Oil Company 1967-68 11 13,116 


Pangea Explorations, Inc. 1987-1991 15 2,110 

Arimetco International, Inc. Unknown Unknown Unknown 

Total ~342 ~78,139 

1.3 GEOLOGY AND MINERALIZATION 

The MacArthur property is one of several copper deposits and prospects located near the town of 

Yerington that collectively comprise the Yerington Mining District. The property is underlain by 

Middle Jurassic granodiorite and quartz monzonite intruded by west-northwesterly-trending, 

moderate to steeply north-dipping quartz porphyry dike swarms. These dikes host a large portion 

of the primary copper mineralization at the nearby Yerington mine and are associated with all 

porphyry copper occurrences in the district. 

The MacArthur copper deposit consists of a 50-150 foot thick, tabular zone of secondary copper 

(oxides and/or chalcocite) covering an area of approximately two square miles. This mineralized 

zone has yet to be fully delineated and remains open to the west and north. Limited drilling has 

also intersected underlying primary copper mineralization open to the north, but only partially 

tested to the west and east. 

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Oxide copper mineralization is most abundant and particularly well exposed in the walls of the 

MacArthur pit. The most common copper mineral is chrysocolla; also present is black copper 

wad (neotocite) and trace cuprite and tenorite. The flat-lying zones of oxide copper mirror 

topography, exhibit strong fracture control and range in thickness from 50 to 100 feet. Secondary 

chalcocite mineralization forms a blanket up to 50 feet or more in thickness that is mixed with 

and underlies the oxide copper. Primary chalcopyrite mineralization has been intersected in 

several locations mixed with and below the chalcocite. The extent of the primary copper is 

unknown as many of the drill holes bottomed at 400 feet or less. The primary copper is currently 

not included in the mine plan for the PEA. 

The MacArthur deposit is part of a large, partially defined porphyry copper system that has been 

complicated by complex faulting and post-mineral tilting. Events leading to the current geometry 

and distribution of known mineralization include 1) Middle Jurassic emplacement of primary 

porphyry copper mineralization by quartz monzonite dikes intruding the Yerington Batholith; 2) 

Late Tertiary westward tilting of the porphyry deposit 60-90° by Basin and Range extensional 

faulting; 3) secondary (supergene) enrichment resulting in the formation of a widespread, tabular 

zone of secondary chalcocite mineralization below outcrops of oxidized rocks called leached 

cap; 4) oxidation of outcropping and near-surface parts of this chalcocite blanket, as well as 

oxidation of the primary porphyry sulfide system. 

1.3.1 Geophysics 

Quaterra contracted three surveys at the MacArthur Project in 2011 and 2012. A borehole 

geophysical survey and a surface IP/resistivity (IPR) survey were carried out by Zonge 

International in 2011, and a detailed helicopter magnetic survey was flown by Geosolutions Pty. 

Ltd. in 2012. These surveys supplement previous geophysical work on the property that includes: 

a 2009 IPR survey carried out by Zonge; a 2007 helicopter magnetic survey carried out by 

EDCON-PRJ; a series of historic aeromagnetic surveys (1966 to 1975) available in analog form 

from the Anaconda Archives; and a series of historic IPR surveys (1963 – 1964) carried out by 

Kennecott Exploration Services/Bear Creek Mining Company and Superior Oil. 

The mineralized system at MacArthur has an anomalous IP and resistivity response first detected 

in the Kennecott and Superior Oil IPR surveys in the 1960’s. The Quaterra 2009 and 2011 IPR 

surveys confirmed the reliability of the earlier surveys and further defined the depth extent of the 

IP anomalies. The 2009 and 2011 Quaterra surveys confirmed that the 1963-64 Kennecott data is 

of good quality and is useful for mapping anomalous IP zones within the upper 1,000-1,200 feet 

from the surface. Below this depth, the older data cannot effectively resolve the bottom of the IP 

anomalies nor determine if any of the anomalies extend to great depths. 

The 2009 and 2011 data sets show this increased depth of exploration is important. Portions of 

the IP response are flat lying with limited depth extent. However both the 2009 and 2011 surveys 

have identified anomalous IP responses with depth extent in excess of 2000 feet and possibly 

feeder zones of the near surface zones. In 2011 two borehole IP surveys were run that 

demonstrate Quaterra’s ability to explore for deep sulfide responses below the depth of 

exploration of surface techniques. The modern data maps subtle low resistivity features which 

are interpreted to be porphyry alteration systems and have identified anomalous IP responses that 

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extend under post-mineral volcanic cover to the north and west of the main MacArthur system. 

These buried anomalies are high priority drill targets. 

Two high resolution helicopter magnetic surveys were flown over the MacArthur Project in 2007 

(EDCON-PRJ) and 2012 (Geosolutions). The modern, high resolution data has a broad 

frequency bandwidth and will be used for 3D modeling and exploring beneath the magnetic 

volcanic cover. 

1.4 EXPLORATION STATUS 

1.4.1 Exploration Drilling Program 

Quaterra has completed 204,656 feet of drilling in 401 holes since beginning drilling in 2007. 

Core holes total 40,233 feet in 58 holes and reverse circulation holes total 164,423 feet in 343 

holes. (Note that one previously listed, but abandoned 115 foot drill hole, has now been removed 

from the database and reported totals). Figure 1-1 show Quaterra's yearly exploration drilling 

footage by year. 

Figure 1-1: Quaterra Exploration Drilling by Year 

Quaterra’s initial objective was to verify and expand the MacArthur oxide resource as had been 

defined by the 1972-1973 Anaconda drilling program and, importantly, to follow up chalcocite 

intercepts in several Anaconda holes as well as in a few outlying early 1960’s holes drilled by 

Bear Creek Mining Company, in late 1960’s drilling by Superior Oil, and holes drilled by the US 

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Bureau of Mines in 1950. Quaterra’s drilling through 2010 successfully expanded the oxide 

mineralization outbound from the MacArthur pit and encountered a widespread, underlying 

tabular blanket of mixed oxide-chalcocite mineralization as well as primary copper intercepts 

that remain incompletely tested. 

During 2011, exploration and infill drilling focused north of the MacArthur pit where earlier 

drilling encountered better grades of oxide and chalcocite mineralization. Holes were angled 

both southerly and northerly to test high angle fractures common in the west-northwest structural 

grain. Strong chalcocite and chalcopyrite mineralization was intersected in several holes in the 

North Ridge zone including QM-183: 1.37% Cu over 40 feet and QM-187: 1.66% Cu over 90’. 

These results were followed by a tightened drill spacing from 500 feet to 250 feet over an 

approximate 2,500 feet by 2,500 feet area north of the MacArthur pit, forming the basis for the 

2011 resource. 

Deep drilling north of the North Ridge Zone intersected significant primary sulfide 

mineralization grading 1.32% Cu over 64 feet in hole QM-164 which is open to the north and 

partially open to the west and east. Although the mineralization at MacArthur has yet to be 

completely closed off to the west and north, the 2011 drilling program expanded and in-filled 

earlier drill results and defined the footprint for the mineral resource estimation published in this 

document. 

1.5 RESOURCE ESTIMATE 

An updated mineral resource estimate has been generated using drill hole sample assays results 

and the interpretation of a geologic model which relates to the spatial distribution of copper in 

the MacArthur deposit. Interpolation characteristics have been defined based on geology, drill 

hole spacing and geostatistical analysis of the data. 

1.5.1 Block Model Definition 

The block model parameters for MacArthur were defined to best reflect both the drill hole 

spacing and current geologic model. Table 1-2 shows the block model parameters used for the 

2011 estimates. 

Table 1-2: MacArthur Model Parameters 

MacArthur East Model Parameters X (Columns) Y (Rows) Z (Levels) 

Origin (lower left corner): 2,429,300 14,685,800 2,800 

Block size (feet) 25 25 20 

Number of Blocks 548 400 150 

Rotation 0 degrees azimuth from North to left boundary 

Composite Length 10 feet (Zone) 

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1.5.2 Assay Database 

An Excel database provided by Quaterra contained the pertinent drill hole and assay information 

for the MacArthur Copper deposit. The database contained 737 drill holes of which 676 drill 

holes from Quaterra and Anaconda (sometimes referred to as the Metech holes) were used. The 

61 holes removed included holes with limited or no information on the assays (Pangea Gold 

1991, Superior, USBM 1952, Anaconda 1955-57), and six Quaterra holes outside the model 

limits. Of the 676 holes used, there are 280 Anaconda (Metech) RC holes and 396 Quaterra holes 

(58 core and 338 RC holes). These drill holes traversed 257,895 feet, producing 51,258 total 

copper sample assay values at a nominal five feet in length. 

A total of 151 drill holes totaling 80,800 feet were added to the database used for the resource 

estimation. These included two holes for which data was unavailable at the time of the last 

estimate, but did not include three 2011 holes which were outside the model limits. 

The variables available in the database are for total copper from Quaterra and Anaconda 

intervals, and acid-soluble copper, a limited number of ferric sulfate soluble (QLT) copper 

assays and a very limited number of cyanide leach copper assays from Quaterra holes. 

1.5.3 Compositing 

The assay data was composited using a 10-foot “zone method”. The zone method is a variant of 

down hole compositing, with the distinction that the composite begins as the drill interval enters 

a rock code zone. This method tends to reduce averaging composites across zones. The process 

first used DataMine ® to assign a MinZone to each 25x25x20-foot block within the model 

specified in Table 1-3. When the majority of a block fell within the interpreted MinZone 

wireframe it was assigned the appropriate code. These coded blocks were then imported into 

MicroModel ® and used to “back-mark” each composite using a simple majority rule. No capping 

was applied. Table 1-3 presents the MinZone codes used in the model. Initial codes of alluvium, 

oxide, oxide and chalcocite mix, and sulfide were 10, 20 and 30 respectively. These codes were 

altered by the addition of 1 if the assays, composites or blocks fell within a 0.12% Cu grade 

envelope predicted by indicator kriging. The codes were also altered by the addition of 100 if the 

data was within a modeled dike. 

Table 1-3: MinZone Codes and Density 

MinZone Code Description Density (cu.ft/ton) 

0 Air and previously mined pit 

Air (0) and Mined 

(12.5) 

5, 6, 105, 106 Alluvium 12.5 

10, 11, 110, 111 Oxide zone 12.5 

20, 21, 120, 121 Chalcocite mix zone 12.5 

30, 31, 130, 131 Sulfide zone 12.5 

9999 Undefined 12.5 

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1.5.4 Geostatistical Analysis and Variography 

A total of twenty-two (21 directional and a omni-directional) variograms were calculated using 

MicroModel® for each MinZone within each area. The program searches along each direction 

for data pairs within a 12.5-degree window angle and 5-feet tolerance band. All experimental 

variograms are inspected so that spatial continuity along a primary, secondary and tertiary 

direction can be modeled. 

Each variogram model was then validated using the “jackknifing” method. This method 

sequentially removes values and then uses the remaining composites to krige the missing value 

using the proposed variogram. 

1.5.5 Kriging and Resource Classification 

Table 1-4 presents the search and kriging parameters employed in the resource model. The 

composite and block codes were used to determine which composites were selected to estimate a 

particular block. 

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Table 1-4: Kriging and Search Parameters 

Tt used a two-part approach to classify the total copper resources. This approach takes into 

account the spatial distribution of the drilling, the distance to the nearest data points used to 

estimate a block, and finally the relative kriging error generated by the estimate. Tt has found 

this approach to be very robust and provide highly reproducible results. The following points 

detail this approach: 

1. A measured block requires 16 samples, with a maximum of five samples per sector in a 

six sector search pattern and a maximum of 2 composites coming from a single drill hole. 

This implies that in most cases, for a block to be classified as measured there must be a 

least 8 drill holes in four cardinal directions. 

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2. The constraints for an indicated block are not as stringent for a measured block. An 

indicated block requires a minimum of 6 samples, with a maximum of 4 samples per 

sector in a sector search pattern and a maximum number of 2 samples coming from a 

single drill hole. This implies that for most cases an indicated block must have at least 3 

drill holes in three of the four cardinal directions. 

3. Relaxing the constraints even more, a inferred block requires a minimum of 2 samples, 

with a minimum of 2 samples per sector in a sector search pattern and a maximum of 2 

composites from a single drill hole. This implies that an inferred block must have a least 

one drill hole from one of the four cardinal directions. 

In addition to the search parameters, kriging error comes into play when determining if a block 

falls into a particular class. Tt has found that by plotting the kriging error as a log-probability 

plot, there is a natural break in the distribution which signifies when the error is too great to 

allow a block to be classified as measured or indicated. In the case of the MacArthur deposit, 

any block with a kriging error of 0.75 or greater was classified as inferred. 

1.5.6 Estimated Resources 

Table 1-5 presents the measured + indicated resources, and Table 1-6 presents the inferred 

resources. The base case cutoff grade for the leachable resource is 0.12% Cu (or TCu) while the 

base case cutoff grade for the primary sulfide resources is 0.15% Cu. Both of these values are 

representative of actual operating cutoff grades in use as of the date of this report. It is Tt’s 

opinion that the MacArthur Mineral Resources meet current CIM definitions for classified 

resources. 

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Oxide and Chalcocite 

Material 

(MinZone 10 and 20) 

Primary Material 

(MinZone 30) 

Table 1-5: Measured + Indicated Copper Resources 

MEASURED+INDICATED COPPER RESOURCES 

MACARTHUR COPPER PROJECT –YERINGTON, NEVADA 

May 2011 

Cutoff Grade Tons Average Grade Contained Copper 

%TCu (x1000) %TCu (lbs x 1000) 

0.50 3,401 0.720 48,974 

0.40 6,730 0.583 78,485 

0.35 10,092 0.513 103,544 

0.30 16,251 0.441 143,171 

0.25 29,859 0.364 217,075 

0.20 65,421 0.286 374,601 

0.18 89,306 0.260 465,106 

0.15 125,659 0.233 585,822 

0.12 159,094 0.212 675,513 

0.50 98 0.720 1,411 

0.40 193 0.586 2,263 

0.35 273 0.523 2,857 

0.30 354 0.478 3,382 

0.25 507 0.416 4,216 

0.20 670 0.369 4,938 

0.18 796 0.340 5,414 

0.15 1,098 0.292 6,408 

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Table 1-6: Inferred Copper Resources 

TABLE 1-6: INFERRED COPPER RESOURCES 


May 2011 


Material 



(MinZone 30) 

1.6 METALLURGY 

Cutoff 

Grade 

%TCu 

Tons 

(x1000) 

Average 

Grade 

%TCu 

Contained 

Copper 

(lbs x 1000) 

0.50 4,294 0.657 56,423 

0.40 9,656 0.538 103,899 

0.35 15,357 0.477 146,444 

0.30 25,851 0.414 213,788 

0.25 43,695 0.356 311,108 

0.20 82,610 0.293 483,929 

0.18 109,920 0.267 587,412 

0.15 166,930 0.232 774,889 

0.12 243,417 0.201 979,510 

0.50 10,644 0.819 174,413 

0.40 18,442 0.653 240,742 

0.35 23,316 0.594 277,181 

0.30 33,831 0.511 345,415 

0.25 53,060 0.423 449,312 

0.20 89,350 0.341 609,188 

0.18 101,375 0.323 654,680 

0.15 134,900 0.283 764,074 

Considering both recent and historical test work, along with information from previous mining 

operations at the MacArthur site, the design basis for this PEA considers a ROM heap leach 

operation with processing of the pregnant leach solution (PLS) through traditional solvent 

extraction / electrowinning (SX/EW). Copper extraction is predicted to range between 60 and 70 

percent depending on material type. Acid consumption projections range between 30 and 35 

pounds of acid per ton of material. The historic MacArthur Pit contains 133 million tons of oxide 

material which is predicted to yield 70% copper extraction with acid consumption of 30 pounds 

of acid per ton of material leached. Material from the MacArthur pit is predominately mined and 

processed over the first 7 years of operation. 

The leach pad will be constructed using an HDPE liner system meeting Nevada requirements 

(NR 455). Conventional solvent extraction will be used. Electrowinning will include permanent 

mother blank stainless steel technology and harvesting of Grade A copper cathode on a 7 day pull 

schedule. All process facilities will incorporate proven industry standard designs and equipment. 

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It is recommended that additional metallurgical test work be performed for the pre-feasibility 

study (PFS), to better understand the metallurgy of this project. A preliminary test program 

design for the PFS is discussed in Section 26 of this PEA. 

The MacArthur Project has a long history of metallurgical testing from 1976 through 2011 

including bottle roll and column leach testing and full scale heap leach operations. Anaconda 

performed the first test work in 1976 and multiple subsequent owners continued test work 

through 2011. The most comprehensive test work was performed by the current owner, Quaterra 

Alaska, Inc., during 2010 and 2011. Quaterra ran a substantial number of bottle roll leach tests 

along with 32 column leach tests on 26 new PQ size core drill holes. These drill holes provided 

reasonable representivity of the MacArthur Project mineral resources. The testwork, both historic 

and that most recently performed, shows the mineralized material is amenable to standard heap 

leaching with good copper extraction. 

The MacArthur Project deposits generally consist of oxidized copper caps transitioning through a 

mixed oxide/secondary sulfide interface into primary sulfides at depth. Of the 271 million tons of 

acid soluble material, 185 million tons is classified as oxide, and 86 million tons is classified as 

mixed or secondary sulfide mineralization. 

Arimetco operated a run of mine (ROM) heap leach/solvent extraction / electrowinning facility 

from 1989 through 1998 leaching low grade oxide stockpile material. Additionally, 6.1 million 

tons of ROM oxide ore from the historic MacArthur pit was leached by Arimetco at the 

Yerington Site. 

1.7 ECONOMIC ASSESSMENT 

The mine and process facilities include a heap leach pad, solvent extraction / electrowinning 

facilities, a sulfuric acid plant with power plant, and the necessary infrastructure to support the 

mine and process facilities. The initial capital cost for the mine and process facilities are 

estimated to be $232.75 million with an additional $147.57 million in sustaining capital. The 

sustaining capital includes a phased expansion of the heap leach pad, additional mine equipment, 

and mobile equipment replacement throughout the life of mine. Closure and reclamation costs at 

the end of the 18 year mine life are estimated to be an additional $82.96 million including a 

salvage value for equipment and materials at mine closure. 

The overall life of mine operating cost for the facilities is $1.89 per pound of recovered copper 

and includes mining, solvent extraction / electrowinning, sulfuric acid plant, general and 

administrative cost, and transportation cost to transport the final cathode copper product to 

market. 

The Net Present Value (NPV) was calculated based on an average annual copper production of 

41.5 million pounds of copper per year and a price of copper of $3.48 per pound. The Project 

will generate after tax NPV of $201.57 million at a discount rate of 8% with an Internal Rate of 

Return of 24.2% and a payback period of 3.1 years. 

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1.8 CONCLUSIONS AND RECOMMENDATIONS 

Quaterra intends to develop the MacArthur Copper Project as a stand-alone project; however, 

other mineral resources owned by Singatse Peak Services (SPS), a subsidiary of Quaterra Alaska 

Inc. at the Yerington mine may be added to the development as appropriate. The MacArthur 

Copper Project has shown potential for development as a large scale copper oxide heap leach 

operation. 

The following additional work is recommended as part of a pre-feasibility study to advance the 

project. 

a) Additional exploration and delineation drilling to better define the resource, particularly 

in the area north of the MacArthur pit and at depth, reduce technical risk and increase the 

project resources. 

b) Update the project resource model with the additional drilling information. 

c) Optimize the mine plan based on the new resource model. 

d) Additional metallurgical test work to confirm the extraction rates and acid consumption 

as outlined in Section 26. 

e) Confirm the design parameters for the heap leach pad, including lift height, irrigation rate 

and inter-lift liners. 

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2 INTRODUCTION 

2.1 GENERAL 

Quaterra Alaska, Inc.’s parent company, Quaterra Resources, Inc. (NYSE Amex: QMM; TSX-V: 

QTA), with headquarters located in Vancouver, British Columbia, Canada, is a mineral 

exploration company focused on making significant base and precious metals discoveries in 

North America. The company also has a local office located in Yerington, Nevada. 

Quaterra requested a number of consultants to provide a Preliminary Economic Assessment 

Technical Report, compliant with Canadian National Instrument 43-101 Standards of Disclosure 

for Mineral Projects, for the MacArthur Copper Project located in Lyon County, Nevada, 

approximately 75 miles southeast of Reno, Nevada. Tetra Tech MM, Inc. of Golden, Colorado, 

was commissioned to prepare an update of the resource estimate and provide a review of the 

metallurgical test work. Independent Mining Consultants, Inc. (IMC) of Tucson, Arizona, was 

commissioned to provide the mining methods and pit design. SRK Consulting (U.S.), Inc. of 

Reno, Nevada, was commissioned to provide the environmental and permitting review; and M3 

Engineering & Technology Corporation of Tucson, Arizona, was commissioned to provide the 

process and infrastructure, capital and operating costs, and the economic assessment for the 

project. 

2.2 PURPOSE OF REPORT 

The purpose of this report is to present updated mineral resource information and a mine 

production plan, process and metallurgical testing information, infrastructure, capital and 

operating costs, a preliminary economic analysis, and other relevant data and information for the 

MacArthur Copper Project since the issuance of the updated mineral resource estimate 43-101 

Technical Report dated January 21, 2011. It is the intent of Quaterra to continue to develop the 

MacArthur Copper Project with possible integration of other resources within the Yerington 

mining district. 

The effective date of this report is May 23, 2012. 

2.3 SOURCES OF INFORMATION 

This report is based on data supplied by Quaterra with the use of historic data from Anaconda, 

Pangea Explorations (Pangea), North Exploration LLC (North), Bear Creek Mining Company, 

The Superior Oil Company (Superior), U.S. Bureau of Mines, and Arimetco International, Inc. 

(Arimetco). Drilling and sampling at the MacArthur site started in 1955 with Anaconda and has 

continued through November 2011 with Quaterra’s last exploration program. 

The information presented, opinions and conclusions stated, and estimates made are based on the 

following information: 

• Source documents used for this report as summarized in Section 27, 

• Assumptions, conditions, and qualifications as set forth in this report, 

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• Data, reports, and opinions from prior owners and third-party entities, and 

• Personal inspection and reviews. 

Tetra Tech, in the preparation of its sections, has not independently conducted any title or other 

searches, but has relied upon Quaterra for information on the status of claims, property title, 

agreements, permit status, and other pertinent conditions. In addition, Tetra Tech has not 

independently conducted any sampling, mining, processing, economic studies, permitting or 

environmental studies on the property. 

Information provided by Quaterra includes: 

• Assumptions, conditions, and qualifications as set forth in the report, 

• Drill hole records, 

• Property history details, 

• Sampling protocol details, 

• Geological and mineralization setting, 

• Data, reports, and opinion from prior owners and third-party entities, and 

• Copper and other assays from original records and reports. 

Additional information provided by third-party entities includes a Preliminary Column Leach 

Study prepared by METCON Research, dated December 2011 and a Scoping Study dated March 

2012 for the Re-mining and Processing of Residual Ore Stockpiles and Tailings at Yerington 

prepared by SRK Consulting. 

2.4 CONSULTANTS AND QUALIFIED PERSONS 

Quaterra contracted a number of consultants, including M3 Engineering & Technology 

Corporation, to provide a review of prior and new work on the project and to prepare technical 

and cost information to support a Preliminary Economic Assessment (PEA) and this Technical 

Report. M3 Engineering & Technology Corporation was responsible for defining the process 

facilities, infrastructure, capital cost, operating cost, preliminary financial assessment, and 

integrating the work by other consultants into a final Technical Report compliant with the 

Canadian National Instrument 43-101 standards. 

Mr. Myron R. Henderson, P.E., of M3 Engineering and Technology Corporation is the principal 

author and Qualified Person responsible for preparation of this report. Mr. Henderson visited the 

site on November 30, 2011 and December 1, 2011 to review the physical conditions and the 

existing infrastructure at site. Other contributing authors and Qualified Persons responsible for 

preparing sections of this report include Dr. Rex C. Bryan of Tetra Tech, Dr. Richard W. Jolk, 

P.E. of Tetra Tech, Mr. Herbert E. Welhener of Independent Mining Consultants, Inc. (IMC), 

and Mr. Mark Willow of SRK Consulting (U.S.) Inc. 

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Dr. Rex C. Bryan, Ph.D., of Tetra Tech is the Qualified Person responsible for preparation of the 

property description, property history, geological setting and mineralization, deposit types, 

exploration, drilling, sample preparation and security, data verification, and description of 

adjacent properties. These sections of the report were taken or updated from the MacArthur 

Copper Project NI 43-101 Technical Report, Lyon County, Nevada, USA dated January 21, 2011 

prepared by Mr. John W. Rozelle, P.G., Principal Geologist of Tetra Tech. Dr. Bryan was also 

responsible for preparation of the updated resource estimate. Dr. Bryan visited the site in 

September 2011 for a physical review of sample preparation and security procedures, as well as 

discussions with geologists and individuals regarding data handling and project geology. It is Dr. 

Bryan’s opinion that there were no deficiencies in the company’s protocols or procedures. 

Mr. Herbert E. Welhener, MMSA-QPM, of IMC was responsible for preparation of the mining 

methods. Mr. Welhener visited the site on November 30, 2011 and December 1, 2011 to inspect 

the physical conditions at the site and the existing MacArthur pit. 

Dr. Richard W. Jolk. P.E., Ph.D., of Tetra Tech was responsible for the review of the new and 

historical metallurgical test work and preparation of the mineral processing and metallurgical 

testing section of this report. Dr. Jolk visited the site on February 20, 2012, March 19, 2012, and 

April 17, 2012. 

Mr. Mark A. Willow, M.Sc., C.E.M. #1832, of SRK Consulting was responsible for the 

preparation of the environmental studies, permitting and social impact section of this report. Mr. 

Willow visited the site on January 30, 2012 and April 17, 2012. 

2.5 DEFINITION OF TERMS USED IN THIS REPORT 

Unless explicitly stated, all units presented in this report are in the Imperial System (i.e. short 

tons, miles, feet, inches, pounds, percent, parts per million, and troy ounces). All monetary 

values are in United States (US) dollars unless otherwise stated. 

Common units of measure and conversion factors used in this report include: 

Linear Measure: 

Area Measure: 

1 inch = 2.54 centimeters 

1 foot = 0.3048 meter 

1 yard = 0.9144 meter 

1 mile = 1.6 kilometers 

1 acre = 0.4047 hectare 

1 square mile = 640 acres = 259 hectares 

Capacity Measure (liquid): 

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Weight: 

1 US gallon = 4 quarts = 3.785 liter 

1 cubic meter per hour = 4.403 US gpm 

1 short ton = 2000 pounds = 0.907 tonne 

1 pound = 16 oz = 0.454 kg 

1 oz (troy) = 31.103486 g 

Analytical Values: 

percent grams per troy ounces per 

metric tonne short ton 

1% 1% 10,000 291.667 

1 gm/tonne 0.0001% 1.0 0.0291667 

1 oz troy/short ton 0.003429% 34.2857 1 

10 ppb 0.00029 

100 ppm 2.917 

Frequently used acronyms and abbreviations: 

ac-ft = acre feet 

ACu or AsCu = Acid Soluble Copper Assay 

Ag = silver 

Au = gold 

Ag oz/t = troy ounces silver per short ton (oz/ton) 

Au oz/t = troy ounces gold per short ton (oz/ton) 

BADCT = Best Available Demonstrated Control Technology 

BLM = Bureau of Land Management 

CIM = Canadian Institute of Mining, Metallurgical, and Petroleum 

CNCu = Cyanide Soluble Copper Assay 

EPA or USEPA = United States Environmental Protection Agency 

EIS = Environmental Impact Statement 

°F = degrees Fahrenheit 

FA = Fire Assay 

ft = foot or feet 

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ft2 = square foot or feet 

ft3 = cubic foot or feet 

GCL = Geosynthetic Clay Liner 

g = gram(s) 

gpl = grams per liter 

gpm = gallons per minute 

h = hour 

HDPE = High Density Polyethylene 

km = kilometer 

kV = kilovolts 

kWh = Kilowatt hour 

kWh/t = Kilowatt hours per ton 

l = liter 

lb(s) = pound(s) 

lbs/ft 3 = pounds per cubic foot 

LME = London Metal Exchange 

MW = megawatts 

NDEP = Nevada Division of Environmental Protection 

NEPA = National Environmental Policy Act 

NSR = net smelter return 

PEA = Preliminary Economic Assessment 

PFS = Preliminary Feasibility Study 

PLS = Pregnant Leach Solution 

PoO = Plan of Operations 

ppm = parts per million 

ppb = parts per billion 

QLT = Quick Leach Test, also Ferric Soluble Copper 

RC = reverse circulation drilling method 

ROD = Record of Decision 

SCFM = standard cubic feet per minute 

SX/EW = Solvent extraction & electrowinning 

TCu = Total Copper Assay 

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ton = short ton(s) 

tph = tons per hour 

tpy = tons per year 

tpm = tons per month 

tpd = tons per day 

tph = tons per hour 

μm = micron(s) 

VLT = Vat Leach Tailings 

% = percent 

Abbreviations of the Periodic Table 

actinium = Ac aluminum = Al americium = Am antimony = Sb argon = Ar 

arsenic = As astatine = At barium = Ba berkelium = Bk beryllium = Be 

bismuth = Bi bohrium = Bh boron = B bromine = Br cadmium = Cd 

calcium = Ca californium = Cf carbon = C cerium = Ce cesium = Cs 

chlorine = Cl chromium = Cr cobalt = Co copper = Cu curium = Cm 

dubnium = Db dysprosium = Dy einsteinum = Es erbium = Er europium = Eu 

fermium = Fm fluorine = F francium = Fr gadolinium = Gd gallium = Ga 

germanium = Ge gold = Au hafnium = Hf hahnium = Hn helium = He 

holmium = Ho hydrogen = H indium = In iodine = I iridium = Ir 

iron = Fe juliotium = Jl krypton = Kr lanthanum = La lawrencium = Lr 

lead = Pb lithium = Li lutetium = Lu magnesium = Mg manganese = Mn 

meltnerium = Mt 

mendelevium = 

Md 

mercury = Hg 

molybdenum = 

Mo 

neodymium = Nd 

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3 RELIANCE ON OTHER EXPERTS 

The MacArthur Copper Project has been an operating mine for several years during the 1990’s 

and has been the subject of numerous written reports. The reports were prepared by mining 

consulting firms on behalf of the owners or operators of the property at the time. M3 Engineering 

and Technology Corporation (M3) have relied on the information provided by Quarterra 

Resources, Inc., its consultants, and the information provided in the numerous reports on the 

Project. The key reports relied on are listed below: 

a) John W. Rozelle, P. G., Principal Geologist, Tetra Tech MM, Inc., “MacArthur Copper 

Project, NI 43-101 Technical Report, Lyon County, Nevada, U.S.A.” dated January 21, 

2011. 

b) Rodrigo R. Carneiro, MS, Director, METCON Research, “MacArthur Project 

Preliminary Column Leach Study (Volume I)” dated December 2011. 

c) J. Pennington, Principal Mining Geologist, Kent Hartley, Senior Mining Engineer, Jim 

Lommen, Associate-Principal Metallurgist, Mark Willow, Principal Environmental 

Scientist; SRK Consulting (USA), Inc.; “Scoping Study for the Re-mining and 

Processing of Residual Ore Stockpiles and Tailings, Yerington Copper Mine, Lyon 

County, Nevada” dated March 14, 2012. 

M3 Engineering & Technology Corporation has relied on Information provided by Dr. Rex C. 

Bryan of Tetra Tech who authored Section 14 – Mineral Resource Estimate. Dr. Bryan also 

reviewed the information from the January 21, 2011, technical report by John W. Rozelle of 

Tetra Tech MM, Inc. regarding Section 4 – Property Description; Section 5 – Accessibility, 

Climate, Local Resources, Infrastructure, and Physiography; Section 6 – History; Section 7 – 

Geological Setting and Mineralization; Section 8 – Deposit Types; Section 9 – Exploration; 

Section 10 – Drilling; Section 11 – Sample Preparation, Analysis and Security; Section 12 – Data 

Verification; and Section 23 – Adjacent Properties. Dr. Bryan is the Qualified Person 

responsible for these sections in the current report. 

M3 has relied on Dr. Richard W. Jolk, P.E. of Tetra Tech MM, Inc. who authored Section 13 – 

Mineral Processing and Metallurgical Testing. M3 reviewed the work performed by Tetra Tech 

and are in agreement with the conclusions reached regarding copper recoveries and acid 

consumption. 

M3 has relied on Mr. Herbert E. Welhener of Independent Mining Consultants, Inc. (IMC) who 

authored Section 16 – Mining Methods and also provided capital and operating costs for the 

mine operation. 

M3 has relied on Mr. Mark A. Willow of SRK Consulting who authored Section 20 – 

Environmental Studies, Permitting and Social or Community Impact. 

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4 PROPERTY DESCRIPTION AND LOCATION 

4.1 LOCATION 



the town of Yerington, Nevada (Figure 4-1 and Figure 4-2). The property is accessible from 

Yerington by approximately five miles of paved roads and two miles of maintained gravel road. 

Topographic coverage is on US Geological Survey “Mason Butte” and “Lincoln Flat” 7.5’ 

topographic quadrangles. The nearest major city is Reno, Nevada approximately 75 miles to the 

northwest. 

4.2 PROPERTY OWNERSHIP 

The property consists of 470 unpatented lode claims totaling approximately 9700 acres on lands 

administered by the US Department of Interior - Bureau of Land Management (BLM) (Figure 

4-3). Sixty one claims are held by Quaterra by means of a mineral lease with option to purchase, 

executed on August 27, 2005, followed by three amendments dated January 16, 2007, August 6, 

2007, and January 9, 2011. The agreement gives Quaterra the right to purchase the claims from 

North Exploration by making three annual payments of $524,000 (option balance) plus interest at 

the rate of six percent per annum by January 15, 2013. Quaterra’s purchase is subject to a two 

percent Net Smelter Return (NSR) royalty with a royalty buy down option of $1,000,000 to 

purchase one percent of the NSR, leaving a perpetual one percent NSR. The agreement with 

North Exploration is in good standing. The remaining 409 claims were staked as lode mining 

claims by Quaterra. These claims are in good standing with all annual payments to the BLM and 

Lyon County having been paid. 

A portion of the MacArthur claim group is also included in the area referred to as the “Royalty 

Area” in Quaterra Resource's purchase agreement for the acquisition of Arimetco’s Yerington 

properties. Under this agreement, MacArthur claims within this area (as well as the Yerington 

properties) are subject to a two percent NSR production royalty derived from the sales of ores, 

minerals and materials mined and marketed from the property up to $7,500,000. The northernmost 

limit of the Royalty Area is shown in Figure 4-3. 

Quaterra’s claims are located in sections 2 and 3, Township 13 North, Range 24 East; in sections 

10-15, 22-27, and 34-36, Township 14 North, Range 24 East; and in sections 18- 20 and 29-31, 

Township 14 North, Range 25 East, Mount Diablo Base & Meridian. Claim outlines and 

boundaries are displayed on Figure 4-2 and Figure 4-3 and a complete listing of the claims with 

serial numbers is included in Appendix B. 

4.3 MINERAL TENURE AND TITLE 

All claims on the project are unpatented lode-mining claims, and as such require a Federal 

annual maintenance fee of $140 each, due by 12:00 PM (noon) of September 1 of each year. 

Further, each lode claim staked in Nevada requires an Intent to Hold fee of $10.50 each, plus a 

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$4.00 filing fee, due 60 days after September 1 of each year. All fees for all claims within the 

project have been paid in a timely manner and all claims are current. 

All claims were staked by placing a location monument (two- by two-inch wood post) along the 

center line of each claim and two- by two-inch wood posts at all four corners, with all posts 

properly identified in accordance with the rules and regulations of the BLM and the State of 

Nevada. Maximum dimension of unpatented lode claims is 600 feet x 1500 feet. The author 

observed various location monuments and claim corners during the field examination. No legal 

survey of the claims has been undertaken. 

4.4 RELEVANT INFORMATION 

Quaterra’s 2007-2008 core and reverse circulation exploration drilling programs were approved 

by the BLM at the Notice of Intent level supported by posting of a $37,075 bond (File Name: 

NVN-083324, 3809, (NV-033)). Quaterra is currently conducting exploration under a BLM Plan 

of Operations / Environmental Assessment (File name 3809 (NV923Z), BLM Bond Number 

NVB001150) and under Reclamation Permit #0294 with the Nevada Division of Environmental 

Protection. 

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Figure 4-1: General Location Map 

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Figure 4-2: Regional Layout May 

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Figure 4-3: MacArthur Property May 

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5 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE 

AND PHYSIOGRAPHY 

5.1 ACCESSIBILITY 

Access to the property from the town of Yerington is approximately three miles north along US 

Highway ALT 95 to Luzier Lane, then west approximately two miles by pavement to the Mason 

Pass road, an improved county gravel road leading two miles northerly to the property (Figure 

4-2). Property entry is along a 100-foot wide improved gravel mine road that accessed the 

MacArthur open pit copper mine during the 1990s. Beyond the MacArthur pit area are several 

historic two-track dirt roads that provide access throughout the property. 

5.2 CLIMATE 

Elevations on the property range from 4,600 to 5,600 feet as low-rolling to moderately steep 

terrain, sparsely covered by sagebrush interspersed with low profile desert shrubs. There are no 

active streams or springs on the property. All gulches that traverse the property are dry yearround. 

The climate is temperate, characterized by cool winters with temperatures between zero 

and 50 °F and warm to hot summers with temperatures between 50 and 100 °F. Average annual 

precipitation is estimated at three to eight inches per year, with a significant part of this total 

precipitation falling as snow and increasing with elevation. Work can be conducted throughout 

the year with only minor stoppage during winter months due to heavy snowfall or unsafe travel 

conditions when roads are particularly muddy. 

5.3 LOCAL RESOURCES AND INFRASTRUCTURE 

The nearest incorporated town is the agricultural community of Yerington located seven miles to 

the southeast along improved gravel roads and pavement. Formerly an active mining center from 

1953 to 1978 when Anaconda operated the Yerington copper mine and from 1995 to 1997 when 

Arimetco operated the MacArthur oxide copper mine, Yerington now serves as a base for three 

active exploration groups: Quaterra Resources, Inc. (MacArthur and Yerington copper properties 

held by Quaterra Alaska and Singatse Peak Services, respectively), Entrée Gold Inc. (Ann Mason 

copper-molybdenum property), and Nevada Copper Corporation (Pumpkin Hollow Copper 

Project) as displayed on Figure 4-2. Yerington hosts a work force active in, qualified for, and 

familiar with mining operations within a one-hour drive to the property. 

Yerington offers most necessities and amenities including police, hospital, groceries, fuel, 

regional airport, hardware, and other necessary items. A propane-fired 220 megawatt electrical 

generating power plant, operated by NV Energy, is located approximately 12 road miles north of 

Yerington accessed off State Highway 95A. The Wabuska railhead is located approximately ten 

miles north of Yerington along State Highway 95A, two miles north of the turnoff to the power 

plant. Drilling supplies and assay laboratories can be found in Reno, a 1.5-hour drive from 

Yerington. Reverse circulation drilling contractors are found in Silver Springs, Nevada 33 miles 

north of Yerington and in Winnemucca and Elko, Nevada areas, from three to five driving hours 

from Yerington. 

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During the Arimetco operating period, approximately 6.1 million tons of leach ore mined from 

the MacArthur pit was trucked approximately five miles south to the former Anaconda 

Yerington mine site onto leach pads (with approved liners). Leach pad sites and ancillary 

facilities for the MacArthur Project are proposed on unpatented claims controlled by Quaterra 

located northeast of the MacArthur pit, as discussed in Section 18 of this report (Figure 18-1). 

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6 HISTORY 

6.1 PROPERTY HISTORY 

Following the early 1860’s bonanza silver discoveries along the Comstock Lode in the Virginia 

City mining district, prospectors stepped out 30 miles to the southeast to investigate the colorful 

oxide copper showings along the Singatse Range within the present-day Yerington mining 

district (Figure 6-1). A majority of the early work (earliest recorded date of 1883) concentrated 

on contact-metamorphic replacement copper deposits hosted in limestone or limey sedimentary 

rocks clustered from four to six miles south-southwest of the MacArthur property (Moore, 1969). 

These contact copper deposits were mined on a small scale, shipping 2,000 to 1.7 million tons of 

copper ore. Most of this early activity took place before and during World War I. Tingley, et al 

(1993) estimates production from the Yerington district at over 85 million pounds of copper from 

1905 to 1920, ostensibly with very little contribution from the shallow prospects in the 

MacArthur area. 

After the 1920s, only minor copper production is recorded from the contact replacement 

prospects and mines (Moore, 1969). The largest nearby operation, located in the Buckskin 

mining district approximately five miles northwest of the MacArthur property, was the 

Minnesota Mine where copper was mined in the early 1920s, but sizeable production of skarn 

(contact) magnetite iron ore began in 1952 with approximately four million tons of ore produced 

by the end of 1966. 

During the 1940s, Anaconda geologists investigated copper showings over the MacArthur 

property and conducted pre-development drilling over the present day Yerington Mine. US 

Government-funded strategic minerals exploration in the early 1950s supported Anaconda’s 

initial development of the Yerington mine (fully funded by Anaconda following expiration of 

strategic minerals funding in the late 1950s). During 1953 to 1978, Anaconda produced 162 

million tons of 0.55% Cu ore amounting to over 1.75 billion pounds of copper from a single 

open pit mine known as the Yerington Mine located five miles south of the MacArthur property 

(Tingley, et al, 1993). Oxide and sulfide copper ores, hosted in a Middle Jurassic porphyry 

system of granodiorite, quartz monzonite, and quartz monzonite porphyry dike swarms, were 

extracted from the Yerington Mine. 

Anaconda, the US Bureau of Mines, Bear Creek Mining Company, Superior Oil and others 

conducted mineral exploration campaigns at the MacArthur property from the mid-1940s 

through the early 1970s. The most significant program was conducted in 1972 to 1973 by 

Anaconda following an extensive trenching and drilling program that resulted in a non-NI 43- 

101 compliant 13 million tons of plus 0.4% Cu mineralization (Heatwole, 1978). 

During the late 1980s, Arimetco permitted heap leaching sites at the Yerington mine site with 

feed sourced from Yerington mine dumps, oxide stockpiles, and vat leach tailings. Arimetco 

expanded their operations to include approximately 6.1 million tons grading about 0.30% Cu 

mined from 1995 to 1997 from what is now the present day MacArthur pit. Based on 1972 and 

1973 Anaconda drilling, Arimetco published a non-NI 43-101 compliant reserve of 29 million 

tons of 0.28% Cu ore remaining in the planned MacArthur pit (MineMarket.com, 2000). 

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Figure 6-1: Major Physiographic Features 

Major Physiographic Features 

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6.2 HISTORICAL RESOURCES 

Anaconda, the US Bureau of Mines, Bear Creek Mining Company, Superior Oil and others 

conducted mineral exploration campaigns at the MacArthur property from the mid-1940s 

through the early 1970s. The most significant program was conducted in 1972 to 1973 by 

Anaconda following an extensive trenching and drilling program that resulted in a non-NI 43- 

101 compliant estimate of 13 million tons of plus 0.4% Cu mineralization (Heatwole, 1978). 

During the late 1980s, Arimetco permitted heap leaching sites at the Yerington mine site with 

feed sourced from Yerington mine dumps, oxide stockpiles, and vat leach tailings. Arimetco 

expanded their operations to include approximately 6.1 million tons grading about 0.30% Cu 

mined from 1995 to 1997 from what is now the present day MacArthur pit. Based on 1972 and 

1973 Anaconda drilling, Arimetco published a non-NI 43-101 compliant reserve of 29 million 

tons of 0.28% Cu ore remaining in the planned MacArthur pit (MineMarket.com, 2000). 

6.3 HISTORIC MINING 

The MacArthur Project area has seen limited historic mining activity, and there is no indication 

of any historic, small-scale, artisanal mining activity. The most recent activity occurred between 

1995 and 1997, when Arimetco mined a limited tonnage of surface oxide copper for heap 

leaching at the historic Yerington Mine site. No consistent, large-scale mining has occurred on 

the site. 

6.4 HISTORIC METALLURGICAL TESTWORK AND MINERAL PROCESSING 

The metallurgical testwork performed on material from the MacArthur property is dated and 

focused on leach performance of material typical of what was historically mined from the 

MacArthur pit. Anaconda, Bateman Engineering (Bateman), and Mountain States R&D 

International (Mountain States) have all performed various metallurgical testwork for the 

MacArthur property. 

Anaconda completed bottle roll and vat leaching tests on crushed ore. Anticipated recoveries 

ranged from 82 to 85% of total copper while consuming 4 to 5 pounds acid per pound recovered 

copper. Bateman ran 18 and 24-inch diameter 20-foot high column leach tests on run-of-mine 

ore and achieved 50 to 60% recovery of total copper while consuming 3 to 4 pounds acid per 

pound copper. Mountain States testing consisted of crushed un-treated ore and acid-cured ore 

column leach testing at 1.5 and 2.5 inch sizes. Mountain States estimated recoveries for the untreated 

ore at approximately 70% of soluble copper at a 2.5 inch crushed ore size with only 

slightly better recovery at a 1.5 inch size. Acid consumption was approximately 3 pounds acid 

per pound copper. Recoveries for the acid-cured ore were increased by 5 to 10%, and the 

indicated acid consumption was reduced by approximately 1 pound acid per pound copper. 

Acid-cured ore also leached faster than the un-treated ore, with recovery times going from 30 to 

60 days down to less than 30 days. 

A more detailed discussion concerning the historical metallurgy in light of the current 

metallurgical work can be found in Section 13, Mineral Processing and Metallurgical Testing. 

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7 GEOLOGICAL SETTING AND MINERALIZATION 

7.1 REGIONAL GEOLOGY 

The MacArthur Project area is located within the western Basin and Range Province in Nevada 

on the east side of the Sierra Nevada Mountains. Within the Basin and Range, north trending 

normal faults have down-dropped basins on either side of upland ranges. In a similar setting in 

western Nevada, the Singatse Range and Wassuk Range form the western and eastern 

boundaries, respectively, of Mason Valley. The MacArthur property, in the Yerington mining 

district, is located in the west-central portion of Mason Valley along the eastern slopes of the 

Singatse Range. 

The regional geology is displayed on Figure 7-1 (Proffett and Dilles, 1984). The oldest rocks in 

the Yerington area of Mason Valley consist of an approximate 4,000-foot thick section of Late 

Triassic, intermediate and felsic metavolcanics and lesser sedimentary rocks, the McConnell 

Canyon Formation, associated with volcanic arc development along the North American 

continent during the Mesozoic. 

This sequence is disconformably overlain by a series of Upper Triassic carbonates, clastic 

sediments, and volcaniclastics that are in turn overlain by the Norian (aka Mason Valley) 

Limestone, a massive limestone nearly 1,000 feet thick. During the Upper Triassic – Lower 

Jurassic, a section of limestones, clastic sediments, tuffs, and argillites, in part correlative with 

the Gardnerville Formation, were deposited. The Ludwig Limestone, containing gypsum, 

sandstone, and arkose, overlies the Gardnerville Formation. 

Mesozoic plutonism, possibly related to the igneous activity that formed the Sierra Nevada 

Mountains, followed during the Middle Jurassic with emplacement of the Yerington batholith of 

granodiorite (field name) composition and the Bear quartz monzonite. Mesozoic plutonism, 

emplaced approximately 169 Ma (Proffett and Dilles, 1984), was closely followed by Middle 

Jurassic quartz monzonite porphyry dikes and dike swarms related to the Luhr Hill granite 

porphyry. Andesite and rhyolite dikes represent the final phase of Mesozoic igneous activity. 

Mesozoic rocks were deeply eroded and then overlain by Mid-Tertiary tuffs and lesser 

sedimentary rocks. Coarser grained andesite dikes are tabbed as Tertiary. The entire package was 

subsequently faulted along north-trending, down-to-the-east dipping faults that resulted in 

extension and major westerly tilting. 

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Figure 7-1: Regional Geology 

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7.2 LOCAL GEOLOGY 

The MacArthur Copper Property is one of several copper deposits and occurrences hosted in or 

related to Middle Jurassic intrusive rocks within the Yerington Mining District, Lyon County, 

Nevada. The Yerington area is underlain by early Mesozoic volcanic and sedimentary rocks now 

exposed along uplands in the Singatse Range to the west and the Wassuk Range to the east. 

These Mesozoic rocks were intruded by three Middle Jurassic batholiths, the oldest known as the 

McLeod Hill Quartz Monzodiorite (field map name granodiorite), followed by the Bear Quartz 

Monzonite that comprise the majority of outcropping rocks on the MacArthur property. A finer 

grained phase of the Bear Quartz Monzonite, known as the Border Phase Quartz Monzonite, 

occurs at the contact between the McLeod Hill Quartz Monzonite and the Bear Quartz 

Monzonite. These batholiths were subsequently intruded during the Middle Jurassic by the Luhr 

Hill Granite, the source of quartz monzonitic (or granite) porphyries, consisting of moderately to 

steeply north dipping quartz-biotite-hornblende porphyry dike swarms, responsible for copper 

mineralization, striking west-northwesterly across the MacArthur property as well as across the 

entire mining district. 

The geologic record is absent until the middle Tertiary when basalt and voluminous ash flow 

tuffs were deposited over the Mesozoic rocks. 

During advent of Basin and Range normal faulting, ca 18-17 Ma, this entire package of rocks 

was down-dropped to the east along northerly striking, east dipping, low-angle faults that flatten 

at depth creating an estimated 2.5 miles of west to east dilation-displacement (Proffett and Dilles, 

1984). Such extension rotated the section such that the near vertically-emplaced batholiths were 

tilted westerly to an almost horizontal position. Pre-tilt, flat-lying younger volcanics now crop 

out as steeply west dipping units in the Singatse Range west of the MacArthur property. Easterly 

extension thus created a present-day surface that in plan view actually represents a cross-section 

of the geology. 

7.3 PROPERTY GEOLOGY 

The MacArthur property is underlain by two Middle Jurassic batholiths, granodiorite (McLeod 

Hill Quartz Monzodiorite) intruded by quartz monzonite, (Bear Quartz Monzonite) both of 

which are intruded by Middle Jurassic quartz porphyry hornblende and quartz porphyry biotite 

(hornblende) dikes. The north dipping porphyry dike swarms follow penetrative west-northwest 

and east-west structural fabrics. Narrow (



The quartz monzonite, formal designation as Bear Quartz Monzonite, cropping out along the east 

part of the claim block and underlying the MacArthur pit, is beige to light gray to off white, fine 

to medium grained, hard but well-fractured, with minor textural variants. Megascopic 

constituents include ~30% orthoclase, ~30% plagioclase, ~ 20% quartz, and 5- to 10-percent 

hornblende. In bench walls at the MacArthur Pit, quartz monzonite hosts conspicuous light 

brown limonite banding (averaging 4 to 6 per foot) sub-parallel to the steeply north dipping, 

west-northwest trending quartz porphyry dikes. Along the eastern portions of the property, 

including the eastern third of the MacArthur pit, quartz monzonite assumes a light gray color due 

to widespread sodic-calcic alteration. 

A phase known as the “border-phase quartz monzonite” is found at the top of the Bear Quartz 

Monzonite pluton (Proffett and Dilles, 1984) and is often mapped at the contact between the 

granodiorite and the quartz monzonite. The border-phase is finer-grained than the quartz 

monzonite and contains more abundant potassium feldspar. 

Quartz-hornblende / biotite porphyry dikes, originating from the Jurassic Luhr Hill Granite 

intrude both granodiorite and quartz monzonite at the MacArthur property and are recognized in 

dike swarms regionally throughout the Yerington mining district. Porphyry dikes hosted a large 

portion of the primary copper mineralization at Anaconda’s Yerington mine and are associated 

with all copper occurrences in the district. Not all porphyry dikes host copper mineralization, be 

it sulfide or oxide. At the MacArthur property, porphyry dikes strike west-northwesterly, dipping 

moderate to steeply north, typically as ridge-formers with widths to 50 feet or more. Porphyry 

dikes at MacArthur are classified by dominant mafic minerals as quartz biotite porphyry and 

quartz hornblende porphyry, each subdivided further based on composition and alteration. Dikes 

contain feldspar crystals and either hornblende or biotite crystals set in an aphanitic matrix. 

MacArthur pit walls offer excellent exposures of the dikes that host (fracture-controlled) oxide 

copper mineralization. The following descriptions originate from Quaterra’s surface mapping 

and from core and chip logging: 

• Quartz biotite porphyry: contains 2 to 4 mm, generally euhedral, blackish biotite “books” 

(5 to 10%) and 2 to 8 mm cloudy quartz phenocrysts (“quartz eyes”) 2 to 5%. 

Hornblende is rare to absent. Feldspars commonly 3 to 5 mm. May host sulfide or oxide 

copper. May or may not have indigenous limonite. If hornblende is present and altered to 

secondary biotite, the dike is mapped as QMpb-2, otherwise mapped as QMpb-1. 

• Quartz hornblende porphyry: contains acicular hornblende crystals, typically thin, 

“needle-like” to 5 mm long; feldspars vary from 2 to 5 mm. Variety QMph-1 contains 1- 

5% sulfide (mostly pyrite) with or without indigenous limonite and 3-5% quartz 

phenocrysts (2 to 5 mm). Variety QMph-2 contains 2-3% sulfides (common) and always 

has indigenous glass (resinous) limonite derived from primary oxidized chalcopyrite, it 

also contains oxide copper, and quartz phenocrysts (2-5 mm) present to 2-5%. 

Variety QMph-3 commonly contains large (to 10 mm) epidote “splotches” (phenocrysts 

or “epidotization”) with 0% to trace fine grained (~1 mm) quartz phenocrysts, 0% to trace 

sulfides. Any oxide copper is transported from nearby copper-bearing rocks and not 

oxidized from the porphyry itself. 

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The best exposures of Jurassic age andesite dikes are found in the walls of the MacArthur Pit 

where the typically soft- to medium-hard, recessive, olive-greenish dikes can be traced from 

bench to bench and in some cases followed across the pit floors. Andesite dikes are commonly 

very fine grained, plagioclase-bearing porphyries that pinch and swell as they fill fractures. Fistsized 

pillows may be a weathering product. Andesite dikes intrude the hornblende and biotite 

quartz porphyry dikes, again best exposed in MacArthur pit walls. Andesite dikes commonly 

contain oxide copper derived from nearby copper-bearing rocks rather than from the andesite 

dikes themselves. 

Jurassic age rhyolite dikes are also well exposed within the MacArthur Pit walls. The rhyolite is 

a white to gray, dense, siliceous rock. Rhyolite dikes contain approximately 5% mafic minerals 

(hornblende and biotite) and rare (1-2%) quartz phenocrysts. Within the MacArthur pit the 

rhyolite can contain oxide copper mineralization; elsewhere on the property it is barren. 

Tertiary hornblende andesite dikes have also been identified on the MacArthur property. These 

dikes are similar, but coarser grained than the Jurassic andesite dikes, containing abundant, 

acicular, black hornblende phenocrysts and occasionally plagioclase phenocrysts up to 5-10 mm 

in long dimension. Tertiary hornblende andesite dikes are frequently observed intruding Basin & 

Range fault structures. These dikes occasionally contain exotic oxide copper mineralization. 

The Mesozoic intrusive rocks are unconformably overlain by a series of nine, moderate to 

steeply west dipping Mid-Tertiary ash flow tuff units with minor mafic flows and tuffaceous 

sediments dated at 27.1 to 25.1 Ma (Proffett and Proffett, 1976). The volcanic units make up the 

uplands in this part and throughout the Singatse Range and cover alteration and structure in the 

Jurassic igneous rocks. 

The dominant west-northwest (N60⁰W to N80⁰W) structural fabric recognized throughout the 

Yerington District is manifested at the MacArthur property as porphyry dike swarms and as high 

angle shears, faults, and joints along which andesite dikes developed. Structure played a key role 

in localizing copper oxide mineralization around the historic pit area, principally along the westnorthwest 

fabric and, secondarily, along generally orthogonal northeast structure bearing N20°E 

to N40°E. 

The MacArthur fault, a low angle, easterly striking, north dipping, normal fault is the largest 

structure recognized on Quaterra’s claims. The hanging wall of the fault displaces the basal unit 

of the Tertiary ash flow tuff sequence approximately 2,000 feet to the east. The displacement of 

Jurassic intrusive as defined by the offset of the contact of the border quartz monzonite with 

granodiorite is on the order of 4,000 feet to the east. The MacArthur fault is one of few faults in 

the Yerington district known to have been active in both Jurassic and Tertiary time. 

Chalcocite/oxide mineralization has a close spatial relation to the trace of the MacArthur fault 

north and west of the MacArthur pit. Gouge in the fault frequently contains chalcocite and/or 

copper oxide suggesting a structural mineralizing trap. 

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7.3.1 Alteration 

Alteration types recognized at the MacArthur property represent those found in mineralized 

porphyry copper systems. A generalized distribution of the MacArthur alteration types is 

displayed in Figure 7-2. The following descriptions are derived from field observation and from 

drill core and chip logging. 

7.3.1.1 Propylitic Alteration 

Propylitic alteration is common throughout the MacArthur property in the granodiorite, quartz 

monzonite, quartz monzonite porphyries, and in the Jurassic andesite. This alteration type occurs 

as chlorite replacing hornblende, and especially epidotization as veining, coatings, and or 

flooding on the granodiorite. Calcite veining is present but not common, observed largely in core 

or drill cuttings. Feldspars are commonly unaltered. Propylitic alteration frequently overprints or 

occurs with the alteration types described below. 

7.3.1.2 Quartz-Sericite-Pyrite (QSP) or Phyllic Alteration 

Phyllic alteration is most frequently characterized by tan or light green sericite partially or 

completely replacing hornblende and/or biotite sites. When phyllic alteration becomes more 

intense, plagioclase and/or K-feldspar sites are also replaced by sericite. Maroon limonite, 

hematite, and trace sulfide (chalcocite, pyrite, and chalcopyrite) accompany sericite. However, 

these minerals do not replace mafic or felsic sites. Sericitic altered zones are often quite 

siliceous; however, it is unclear if it is due to quartz addition or simply the destruction of other 

primary minerals. 

Phyllic alteration is most pervasive and intense in the Gallagher area and in the northeastern part 

of the deposit, around hole QM-072. Weak and less pervasive phyllic alteration is found just 

west of the MacArthur pit and in limited areas around the MacArthur fault. The alteration type 

does not show preference with rock type and has been described in the granodiorite, quartz 

monzonite, and quartz monzonite porphyries. 

7.3.1.3 Potassic Alteration 

Potassic alteration occurs as shreddy, fine-grained biotite replacing hornblende and rarely as 

pinkish potassium feldspar flooding or in vein haloes, along with disseminated magnetite. 

Potassic alteration as shreddy secondary biotite is most obvious in the western and central areas 

of the MacArthur pit. However, there is occasional biotite replacing hornblende in the 

northwestern and western portions of the MacArthur property, but is usually less than 20%. Kfeldspathization 

is conspicuous at the base of the mineralized drill intercept of drill hole QM- 

100. Potassic alteration of some degree has been identified in the granodiorite, quartz monzonite, 

and quartz monzonite porphyries. 

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7.3.1.4 Sodic-Calcic Alteration 

Pervasive sodic-calcic alteration has been identified within the eastern portions of the MacArthur 

pit and as broad zones in the far northeastern portion of the district and south of the MacArthur 

pit. This type of alteration most frequently occurs as albite replacing K-feldspar and as chlorite 

replacing hornblende in the quartz monzonite, Sodic-calcic alteration has also been identified in 

the granodiorite and quartz monzonite porphyries. Epidote staining and phenocrysts as well as 

sphene crystals are ubiquitous. Actinolite replaces hornblende in the more intense zones of sodiccalcic 

alteration occurring most commonly in the Albite Hills east of the MacArthur pit. 

7.3.1.5 Silicification 

Silicification occurs as a wholesale replacement of the rock, but only occurs as small and 

irregular zones that are less than 200 feet across. Typically silification is confined as a narrow 

halo (less than five feet) along structure and quartz veining. Silicification is present in the 

western portion of the district, around the Gallagher area and as isolated occurrences within the 

MacArthur pit. 

7.3.1.6 Multiple alteration types 

Multiple alteration types are common throughout the area and tend to occur together. Shreddy 

chlorite has been identified in the MacArthur pit, which likely represents propylitic alteration 

overprinting potassic alteration. Zones of QSP and propylitic alteration have been identified 

between the Gallagher area and the MacArthur pit. 

7.3.1.7 Supergene alteration 

Sulfuric acid (H2SO4), formed by the oxidation of sufides, has altered feldspars and mafic 

minerals to clay and sericite. At the Gallagher area and north of the MacArthur pit, supergene 

alteration has formed leached capping which is underlain by chalcocite mineralization. 

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7.4 MINERALIZATION 

Figure 7-2: Generalized Alteration Types 

Copper mineralization has been identified across nearly the entire area investigated by Quaterra’s 

drilling program at MacArthur and gives every indication of extending well beyond. As currently 

defined by drilling, copper mineralization covers an area of approximately two square miles as 

defined by drill holes on 500 feet to 250 feet spacing north of the MacArthur pit to 

approximately 150 feet spacing within the pit. 

Oxide, chalcocite, and primary copper mineralization is hosted in both granodiorite and quartz 

monzonite, and in quartz biotite-hornblende (quartz monzonite) porphyry dikes all of middle 

Jurassic age. An insignificant percentage of oxide copper is also hosted in northwest striking 

andesite dikes that make up less than approximately one to two percent of the host rocks on the 

property. Fracturing and favorable ground preparation supplied the passage ways for the copper 

to migrate. 

Copper oxide minerals are exposed throughout Quaterra’s MacArthur property, in MacArthur pit 

walls as primarily green and greenish-blue chrysocolla CuSiO3·2H20 along with black neotocite, 

aka copper wad (Cu, Fe, Mn) SiO2, with very minor azurite Cu3(OH2)(CO3) and malachite 

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Cu2(OH2)CO3, while tenorite (CuO) was identified with the electron microprobe (Schmidt, 

1996). Copper-enriched limonite was identified by Anaconda as the mineral delafossite 

(CuFeO2). Chalcocite has been identified in drill holes below and north of the MacArthur pit and 

in drilling throughout the property. The sulfides digenite (Cu9S5) and covellite (CuS) have been 

identified petrographically in drill cuttings. Bornite (Cu5FeS4) has also been identified 

petrographically in the Gallagher area. The oxide copper mineralization is fracture controlled, 

coating joint and fracture surfaces and within shears and faults. Both green and black copper 

oxides are frequently found on 1-5 millimeter fractures, as coatings and selvages and may be 

mixed with limonite. The fractures trend overall N60°W to N80°W (bearing 300° to 280° 

azimuth) and generally dip to the north. Limited turquoise is found on the property, mainly in 

small veinlets. On a minor scale, oxide copper mineralization replaces feldspar phenocrysts in 

the igneous host units, favoring andesite. 

A significant amount of chalcocite has been intersected in drill holes. Chalcocite is seen on drill 

chips or drill core coating pyrite and replacing chalcopyrite as tiny, blackish “dustings” and thin 

to thick coatings, strongest when occurring on and near the MacArthur fault. Chalcopyrite is 

present as disseminations and veinlets, with or without chalcocite. As much of the historic and 

current drilling was stopped at shallow (



8 DEPOSIT TYPES 

The MacArthur deposit is a supergene enriched, oxidized porphyry copper system. Although the 

porphyry system likely developed in near-vertical geometry, regional studies by Proffett and 

Dilles (1984) suggest the MacArthur area is tilted westerly approximately 60 to 90 degrees from 

its original vertical position and extended to the east so that the map view is actually a structural 

cross section. The original northwest strike of the near vertical porphyry dikes resulted in a 

northerly dip of the structures with the post mineral tilting (Figure 8-1). 

Figure 8-1: Datamine© View of Resource Block Model Looking West 

(Figure 8-1 View of Datamine© resource block model with planned open pits looking West below the North Ridge 

showing the northerly dip of primary sulfide mineralization (red).) 

The alteration visible in outcrops and drill samples is consistent with the west tilted, near 

horizontal orientation of the porphyry system. Phyllic alteration from the upper portion of the 

porphyry system dominates to the west. The alteration grades to potassic in the central 

MacArthur pit area and pervasive sodic-calcic alteration dominates in the eastern portions of the 

MacArthur pit and in the far northeastern portion of the district. 

Copper occurrences in the MacArthur pit area are related to primary copper sulfides associated 

with the porphyry copper center. The primary chalcopyrite (CuFeS2) was enriched by supergene 

chalcocite (Cu2S) and later exposed to oxidation forming chrysocolla (CuSiO3) and black 

copper wad (Cu,Fe,Mn SiO2). In the North Ridge area the chalcocite blanket shows only minor 

oxidation. The supergene blanket follows current topography except to the north of 14,691,501E 

(approximately) where it has a shallow dip to the north (Figure 8-2 and Figure 8-3). 

Primary porphyry copper sulfides have also been intersected north of the North Ridge area in 

drill thicknesses up to 100 feet and in the Gallagher area. These intercepts maybe related to the 

MacArthur pit porphyry center or a new, yet to be discovered porphyry copper deposit. 

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Figure 8-2: East-West Section 14,691,000N (Looking North) 

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Figure 8-3: North- South Section 2,438,324 (Looking West) 

Figure 8-3 

Drillhole Section 2,438,324E 

(Looking West) 

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Fig8-3.cdr



Exploration 

Starting in April 2007 and continuing through October 2008, and from December 2009 through 

November 2011 Quaterra completed extensive reverse circulation and core drilling at the 

MacArthur property. Drill results through October 2008 coupled with 1972-1973 Anaconda 

drilling provided the data for Tetra Tech to publish the February 2009, revised March 2009, 

MacArthur NI43-101 Technical Report. An additional 77 drill holes completed through 

September 2010 formed the basis for the January 2011 NI 43-101 Technical Report. During 

2011 an additional 152 holes were completed, and are the basis of this updated Technical Report. 

There are three different mineralization zones encountered at MacArthur. All three 

mineralization zones - oxide, mixed chalcocite/oxide, and primary sulfide - have grown with 

additional drilling and none are yet entirely closed off. 

8.1 OXIDE ZONE EXPLORATION 

Extents of the oxide mineralization on the property remain open to the west and are only partially 

defined to the south. 

Five thousand feet west of the MacArthur pit, Quaterra holes QM-133 and QM-153 intersected 

0.27% Cu over 235 feet and 0.16% Cu over 125 feet, respectively, of oxide and acid soluble 

copper. The mineralization is open 1,000 feet farther to the west. 

Southeast of the MacArthur pit, holes spaced from 500 to 1,000 feet apart contain 0.1 to 0.3% Cu 

intercepts. Drill holes QM-142, QM-108, and QM-140 encountered 0.21% Cu over 50 feet to 

0.31% Cu over 10 feet in an area that remains untested for 3000 feet to the Shuman area (3,500 

feet south of the MacArthur pit) where oxide intercepts of 0.24% Cu over 45 feet and 0.39% Cu 

over 30 feet were encountered from surface. Mineralization in these holes (referred to as the 

Shuman drill holes) is open in all directions, but obscured to the south by Tertiary volcanic 

cover. 

8.2 CHALCOCITE/OXIDE ZONE EXPLORATION 

Chalcocite/oxide mineralization remains open to various degrees in all directions, in light of the 

500 foot drill spacing. Chalcocite mineralization is partially open to the northwest. 

8.3 PRIMARY SULFIDE ZONE EXPLORATION 

Primary, porphyry-style copper mineralization has been encountered at the North Porphyry 

Target area and is described in the following paragraph. In the Gallagher area, primary copper 

mineralization occurs from 450 feet depth in QM-10, with 0.43% Cu over 155 feet to 0.74% over 

76 feet in QM-46 from 1,279 feet depth as chalcopyrite disseminations and veinlets. Additional 

drilling to target primary sulfide mineralization is warranted for as there are only eight holes 

exceeding 800 feet depth over an approximate one half square mile area. 

Quaterra’s drilling program at the North Porphyry Target, some 3,000 feet north of the 

MacArthur pit, encountered 115 feet of mineralization (partially enriched with chalcocite) 

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averaging 1.15% Cu at a depth of 470 feet in drill hole QM-68. A similar section of 

mineralization in QM-070 (500 feet east of QM-068) averaged 1.02% Cu over a thickness of 45 

feet at a depth of 435 feet. Together with mineralized intercepts in QM-072, (500 feet east of 

QM-070) which cut 15 feet of 1.2% Cu, the results indicated a possible porphyry center in the 

foot wall of the MacArthur fault. In 2010 this concept was favorably tested 1,500 feet north of 

QM-68 where drill hole QM-100 intersected 0.58% Cu over 65’ from 1203.5 feet. During 2011, 

QM-100 was offset 1,000 feet north by QM-164 returning 1.32% Cu over 64 feet from 1,673 feet 

depth. These primary sulfide intercepts define a 6,000 foot mineralized zone (corridor), including 

the oxide mineralization at the MacArthur Pit north to the sulfide intercepts in QM-164, untested 

500 feet east and west of QM-100 and QM-164 and open to the north. 

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9 EXPLORATION 

9.1 GEOPHYSICS 

Quaterra contracted three surveys at the MacArthur Project in 2011 and 2012. A borehole 

geophysical survey and a surface Induced Polarization/Resistivity (IPR) survey were carried out 

by Zonge International in 2011. A detailed helicopter magnetic survey was flown by 

Geosolutions Pty. Ltd. in 2012. These surveys supplement previous geophysical work on the 

property that includes: a 2009 IPR survey carried out by Zonge; a 2007 helicopter magnetic 

survey carried out by EDCON-PRJ; a series of historic aeromagnetic surveys (1966 to 1975) 

available in analog form from the Anaconda Archives; and a series of historic IPR surveys (1963 

– 1964) carried out by Kennecott Exploration Services/Bear Creek Mining Company and 

Superior Oil. 

The 2009 and 2011 IPR surveys were designed to confirm the reliability of the earlier surveys 

and to further define the depth extent of the IP anomalies. The 1963-64 data indicate that the 

zone of anomalous IP response is typically flat-lying with a thickness of less than 1,000 feet and 

does NOT extend beyond a depth of 1500 feet. Comparison of data from the surveys done more 

than 45 years apart is that the 1963-64 Kennecott data is of good quality and can be used 

effectively to define IP anomalous zones within the upper 1,000-1,200 feet of the subsurface. 

However beyond that depth the 1963-64 data cannot effectively resolve the bottom of the IP 

anomalies nor determine if any of the anomalies extend to great depths. The modern data sets 

show this increased depth of exploration is important. For example a portion of the anomalous IP 

zone on line 306075E is depth limited however the anomalies on the north and south ends of the 

lines extend much deeper, to a depth exceeding 2,000 feet. 

The 2007 EDCON-PRJ high-resolution, helicopter-borne aeromagnetic survey was flown over 

the MacArthur Copper Project. The survey was designed so that data from historic Anaconda 

surveys (1966 to 1975) could be merged with the new data. The historic surveys were recovered 

from the Anaconda Archive collection maintained by the American Heritage Center, University 

of Wyoming. EDCON-PRJ digitized the historic survey data from the paper maps, as no digital 

data was available for those surveys. 

Note that all modern geophysical surveys have been run on Nad27 UTM Zone 1N metric grids, 

but for purposes of consistency, depths and distances are given in feet. 

9.1.1 IP/Resistivity Surveys 

9.1.1.1 2011 Work 

Surface and down hole IP/resistivity (IPR) surveys were run at the MacArthur Project area in 

2011. The surface IPR survey was conducted by Zonge International in February of 2011. The 

primary purpose was to continue to cover the project area with modern high quality IPR data to 

replace the historic data collect by Kennecott and Superior Oil in the early 1960’s. The goal of 

the survey was to map sulfide and alteration response at depth within the central alteration zone 

(Figure 9-1) and beneath volcanic cover adjacent to the alteration zone. Four lines of surface IPR 

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were run in 2011. The quality of the data recorded is good. Previous interpretations are supported 

and a number of new targets have been identified. Of particular interest are targets identified 

beneath volcanic cover to the north and west of the main alteration zone and low resistivity/high 

IP phase anomalies which continue to depth indicating possible fluid feeder zones from depth. 

Figure 9-2 through Figure 9-7 show the pseudo-sections and inversion models for three of the 

2011 IPR lines. In each figure the top panel shows the inversion model for IP or resistivity. The 

observed data is shown in the middle panel and the bottom panel is the calculated pseudo-section 

generated from the inversion model. 

The IP models and pseudo-sections for lines 4300 (Figure 9-2) and 4900 (Figure 9-4) run over 

the Gallagher Zone and continue into the volcanic cover to the north. Both lines show deeper IP 

response continuing under the near surface volcanic cover (see black arrow). Although the 

amplitude of the deeper response is lower, 20 to 30 milliradian (mrads) at depth versus up to 80 

mrads at the surface this may not accurately reflect sulfide content at depth. Also note that the 

base of the IP response is poorly resolved and there is some evidence of continuation to depth on 

all four lines (L4300, L4900, L5300 and L7500) runs in 2011. Figure 9-3 and Figure 9-5 on line 

L4300 and L4900 respectively show lower resistivity zones associated with the higher IP 

responses. These low resistivity zones are interpreted to indicate alteration associated with 

mineralization. A possible interpretation of the deeper IP and resistivity anomalies is that they 

are feeder zones for the shallower, flat lying mineralization. 

Figure 9-6 and Figure 9-7 show IPR models and pseudo-sections for line L7500 located to the 

east of the North Porphyry target. The top of the IP response is approximately 800 feet (250 

meters) depth and the response is weak. A low resistivity zone surrounds the IP response. This 

response may indicate a porphyry system style of zonation. 

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Figure 9-1: IPR line locations over the central MacArthur Project area. 

(Figure 9-1: The 2011 lines are shown in black and the 2009 lines in grey. The lines are plotted on an image of the 

Reduced to Pole - Total Magnetic Intensity (RTP) data acquired by EDCON-PRJ in 2007. The magnetic low located 

between the Gallagher Adit and the North Porphyry target is interpreted to represent the central alteration zone 

which is targeted by the IPR data set. The location of drill holes QM-164 and QM-177 used for borehole IP surveys 

are shown and are further discussed within the text). 

The second ground geophysical program carried out at the MacArthur Project in 2011 was a 

down hole IP/resistivity survey utilizing drill holes QM-164 and QM-177. This work was carried 

out by Zonge International in August 2011. The surveys were designed to explore for sulfide 

response at depths greater than can be resolved using surface arrays. The goal was to determine 

which direction from the drill hole sulfides occur. The drill hole acts as the pathway to place an 

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electrode to depth. The technique maps sulfides at or above the level of the buried electrode. 

Drill holes QM-164 and 177 are of moderate depth and were used to test the process for 

emplacing electrodes in holes that are difficult to keep open (Figure 9-8 and Figure 9-9). 

Figure 9-2: Line 4300 (304300E) IP pseudo-section and inverted phase/depth model 

(Figure 9-2 - Line 4300 runs over the western side of the Gallagher target area. The primary feature of interest on 

this line is extension of the near surface phase anomaly to depth (800-1000 feet (200 to 300 meters)) under volcanic 

cover to the north (see black arrow). Approximate location of QM-177(328 feet (100 meters) to east) is shown as 

circled + sign.) 

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Figure 9-3: Line 4300 Resistivity pseudo-section and inverted resistivity/depth model 

(Figure 9-3 - Line 4300 (304300E) The weak low resistivity zone may be associated with alteration coincident with 

the buried sulfide system (see black arrow). 

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Figure 9-4: Line 4900 IP pseudo-section and inverted phase/depth model 

Figure 9-4 - Line 4900 (304900E) runs over the eastern side of the Gallagher target area. Note phase response 

beneath volcanic cover (black arrow). 

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Figure 9-5: Line 4900 Resistivity pseudo-section and inverted resistivity/depth model. 

(Figure 9-5 - Line 4900 (304900E) The weak low resistivity zone is possible alteration coincident with the buried 

sulfide system (black arrow)). 

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Figure 9-6: Line 7500 IP pseudo-section and inverted phase/depth model 

(Figure 9-6 - Line 7500 (307500E) cuts off the MacArthur Pit North Target to the east. The weak phase anomaly 

(20+ mrads) indicates the continuation of sulfides to depth.) 

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Figure 9-7: Line 7500 Resistivity pseudo-section and inverted resistivity/depth model 

(Figure 9-7 - Line 7500 (307500E) The weak resistivity lows surround the IP response and may indicate a classic 

porphyry alteration zonation.) 

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Figure 9-8: QM-164 down hole electrode to remote electrode transmitter pair 

(Figure 9-8 – IP response plotted in upper image and resistivity response in lower image. No significant IP 

response was detected in this drill hole.) 

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Figure 9-9: QM-177 down hole electrode to remote electrode transmitter pair 

(Figure 9-9 - IP response plotted in upper image and resistivity response in lower image. A significant IP response 

is located to the east of the drill hole in this image.) 

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9.1.1.2 2009 and older IP/Resistivity (IPR) Surveys 

Seven lines of surface IPR were run in 2009. These lines together with the 2011 IPR lines make 

up the modern data set which replaces the historic data sets in this area. The purpose of this 

survey was to confirm the results from historic early 1960’s surveys run by Kennecott and 

Superior Oil. Those surveys although useful in initially detecting sulfide response were recorded 

on old generation analog systems and have limited depth extent and unknown quality. 

Figure 9-10 shows the location of the IPR lines conducted in 1963-64 by Kennecott Exploration 

Services (KES) for the Bear Creek Mining Company. KES collected 11 lines of IPR data which 

are plotted in black. The Superior Oil IP lines have not been used in this compilation as the data 

quality and line locations are questionable. 

The 2009 Zonge data confirmed the results of this previous work, explored to greater depth with 

higher quality data and essentially replaced the older data. Seven (7) lines were surveyed and are 

plotted in white (Figure 9-10). 

To put the 2009 and 2011 surveys in perspective it has been observed that high quality IPR 

surveys are capable of sensing and mapping metallic sulfide concentrations of pyrite and/or 

chalcopyrite as low as 1-2% by volume. A significant volume of rock containing 3-5% 

pyrite/chalcopyrite will result in an IP anomaly exceeding 30-40 milliradians, whereas 7-10% 

metallic sulfides will result in anomalies exceeding 75 milliradians. (Nelson and Van Voorhis, 

1983) Both the 2009 and 2011 surveys are of this high quality. 

A number of gaps existed in the 2009 data set which were later filled in by the 2011 survey 

(described above) or are yet to be filled in by future work. 

Figure 9-11 is a summary interpretation of the historic and 2009 IP data sets plotted on a 

magnetic susceptibility inversion image. It is important to note that the stronger amplitude IP 

responses, shown as red bars along the lines, generally reflect shallow responses. The fact that 

the deeper responses are lower amplitude may not reflect relative sulfide content accurately. 

This is important as the exploration program under volcanic cover develops. 

Note that the Central Zone (outlined in brown) is characterized by low magnetic susceptibility, 

probably destruction of magnetite, and high IP effect due to increased sulfides (Figure 9-1 and 

Figure 9-11). Where the NW trending Qmp dike swarm occurs at the SW edge of the Central 

Zone (Figure 9-1) the magnetic susceptibility increases due to magnetite in the dikes. The IP 

effect is high in this area as well. 

The North Porphyry Target, located NNE of the Central Zone (outlined in grey, Figure 9-1 and 

Figure 9-11) is characterized by high IP effect and increased magnetic susceptibility due to 

quartz porphyry dikes. The strongest IP anomalies are coincident with the more intense magnetic 

highs. 

The Gallagher Adit area occurs to the SW of the Central Zone. Similar to the North Porphyry 

target, the northern edge of the Gallagher Adit area is characterized by a zone of moderate 

magnetic susceptibility (Qmp) with zones of moderately strong to strong IP anomalies. 

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Figure 9-10: Line location of the 1960’s Kennecott lines (in black) and the 2009 

replacement line (in white). 

The NW and NW Gallagher Targets are shown by the grey outlines and arrows pointing under 

the Tertiary volcanic cover (Figure 9-1 and Figure 9-11). Alteration and copper mineralization as 

well as zones of coincident high magnetic susceptibility and IP response continue to the contact 

with the post-mineral Tertiary volcanic front in the western portion of Quaterra’s claim block. 

Lines 4300 and 4900 from the 2011 survey indicate those IP responses continue under the 

volcanic cover. 

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Figure 9-11: Historic and 2009 IP data on a modeled magnetic susceptibility depth slice 

(Figure 9-11 A qualitative interpretation of the historic and 2009 IP data plotted on a modeled magnetic 

susceptibility depth slice. Note all of the surface IP work at MacArthur, including the 2011 survey (not shown here) 

cover this central zone between the Gallagher Adit and the North Porphyry target. ) 

Figure 9-12 shows the inverted IP and resistivity model for line 6075 (306075E) recorded in 

2009. The line runs directly over the MacArthur pit to North Porphyry target area. The IP model 

shows a flat lying near surface response with deep responses to the north and south of the Central 

zone. The resistivity model shows the low resistivity alteration pattern associated with the 

modeled IP response. This is a good example of strong IP response surrounded by low resistivity 

due to alteration. The deep, vertical features may be reflecting feeder zones continuing to depth. 

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Figure 9-12: Inversion model and pseudo-sections for line 6075 recorded in 2009. 

Figure 9-12 Line 6075 (306075E) The line runs directly over the MacArthur pit to North Porphyry target area. The 

importance of this line is that it indicates that both the IP response and low resistivity alteration zones are open to 

depth. One interpretation is that these deep features are feeder zones for mineralizing fluids coming from depth. 

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9.1.2 Airborne Magnetic Surveys 

9.1.2.1 2012 Survey 

A small 428 line kilometer airborne magnetic survey was flown over the northern extension of 

the MacArthur Project area in April of 2012 by Geosolution’s Pty Ltd. The block is located to 

the north of the previous survey as shown in Figure 9-13. The flight line direction is N-S, line 

spacing 164 feet (50 meters) and the attempted sensor terrain clearance is 98 feet (30 meters) 

although the mean sensor clearance is somewhat higher, 148 feet (45 meters) due to steep 

topography which required the pilot to fly higher in some areas. 

The objectives of the survey were: 1) to map the contact between batholithic intrusive and 

sedimentary basement; 2) to explore for magnetite rich skarn bodies along this contact; 3) to map 

quartz porphyry dikes and other intrusives; and 4) to map alteration of volcanic rocks, 

destruction of magnetite, associated with porphyry style mineralization in this area. 

Interpretation of this data set is currently in progress and results will not be discussed in this 

report. 

The particular magnetometer system that was used was designed to maintain high frequency 

information that will allow 3D modeling of the broad band magnetic data set. The 3D model will 

be used to explore beneath volcanic cover which masks the magnetic response of deeper units of 

interest. 

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Figure 9-13: Location of the 2012 detailed helicopter magnetic survey 

(Figure 9-13 – Location of the 2012 detailed helicopter magnetic survey (black polygon) with respect to the 

previous magnetic and IP/resistivity work. Note the NE-SW striking ellipse is interpreted to be associated with the 

mineralized alteration zone.) 

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9.1.2.2 2007 Survey 

The 2007 detailed magnetic data set was flown over the MacArthur land block by Edcon-PRJ 

(see Figure 9-1 and Figure 9-13). The data set was flown with a stinger mounted system at a 

terrain clearance of 328 feet (100 meters). The combination of a stinger mounted system and the 

large terrain clearance results in the removal of the high frequency information required in 3D 

modeling. Because of the frequency content difference between the 2007 and 2012 data sets the 

initial modeling program will not use the 2007 data. Ultimately the data sets will be merged and 

modeled together but as a second pass at the modeling. 

The 2007 MacArthur dataset (Figure 9-1) illustrates several features that correlate to the geology, 

alteration and mineralization at MacArthur. The magnetic field in the MacArthur area is 

dominated by intense highs and lows caused by Tertiary volcanic rocks. The northwest quarter 

and the southeast corner of the MacArthur claim block contain highly magnetic volcanic units. 

These areas are denoted in the figure by “Tv”. The intense magnetic lows (deep blue in Figure 

9-1) correlate with specific geologic units within the Tertiary volcanic sequence. Some of these 

units have very strong remnant magnetization which has a major component in the opposite 

direction to the current magnetic field. Hence the strong magnetic field lows. 

The area between the two Tertiary volcanic “fronts” contains the altered and mineralized 

MacArthur hydrothermal system. It appears as a zone of moderately suppressed magnetism but 

not the intense lows associated with remnant magnetization. This zone is approximately 3 miles 

long, NE-SW and 2 miles wide, NW-SE. Alteration, favorable Jurassic dikes, and mineralization 

extend to the edges of Tertiary volcanic rocks, and likely continue under the post-ore ‘volcanic 

cover’ in some areas. 

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10 DRILLING 

10.1 HISTORICAL 

10.2 EXPLORATION & DRILLING HISTORY 

Although the MacArthur area is dotted with numerous shallow pits and prospects, there is little 

available published information. Over the history of the project, several operators have 

contributed to the current drill hole database of more than 300 holes. Table 10-1 summarizes the 

exploration history of the MacArthur area prior to Quaterra’s entry. Figure 10-1 shows the 

location of all historical drill holes. 

Operator 

Table 10-1: Historic Exploration Drilling 


February 2009 

Drill Program 

Date Range 

Number of 

Holes Drilled 

Feet Drilled 

U.S. Bureau of Mines 1947-50 8 3,414 


Bear Creek Mining Company 1963-?? ~14 Unknown 

Superior Oil Company 1967-68 11 13,116 


Pangea Explorations, Inc. 1987-1991 15 2,110 

Arimetco International, Inc. Unknown Unknown Unknown 

Total ~342 ~78,139 

During the late 1940s, Consolidated Copper Mines consolidated various claims into a single 

package that became known as MacArthur, and then attracted the interest of the US Bureau of 

Mines during their investigation and development of domestic mineral resources. The Bureau of 

Mines completed 7,680 feet of trenching in 1948 and followed up with eight diamond drill holes 

for 3,414 feet in 1950 (Matson, 1952). Five of the US Bureau of Mines’ holes (#1-5) fall within 

the northern segment of the present day MacArthur open pit (Table 10-2). Holes #6-8 were 

collared in an area of widespread iron oxide staining approximately 2,000 feet north of the 

MacArthur pit within Quaterra’s mineral resource mine plan footprint. Oxide copper was 

intersected in the southern holes #1-5 while secondary, sooty, chalcocite enrichment was found 

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in the northern holes #6-8. Following the US Bureau of Mines exploration and drilling programs, 

Consolidated Copper abandoned their claims. 

Table 10-2: U.S. Bureau of Mines 1947-1950 Drilling Highlights 

Hole ID Total Depth 

(feet) 


Feb-09 

Key Intercepts 

(Interval or thickness in feet 

and % Cu) 

Notes 

Hole 1 220 110+: 0.2% Bottomed in +0.2% Cu 

Hole 2 556 (-45º) 509-556: 0.55% Bottomed in 0.55% Cu 

Hole 3 428 245-286: 0.40% 

Hole 4 469 (-45º) 79-114: 0.82%, av. 0.2+/-% Lost hole 

Hole 5 510 291+: 0.25%; av.. 0.2+/-% Bottomed in 0.25% Cu 

Hole 6 409 241-303: 0.61%. 303+: ~0.15% 

Hole 7 428 262-297: 0.51% 

Bottomed in 0.2% Cu 

Hole 8 394 250-299: 0.36% Lost hole 

During the middle 1950s, Anaconda, by then operating the Yerington Mine, acquired leases and 

began investigations at MacArthur including 33 shallow drill holes (only 11 exceeding 100 feet) 

during 1955, 1956, and 1957. Six Anaconda holes (#’s 12, 14-17, and 19) fall within the current 

MacArthur pit limits. Key interval assay results from the holes exceeding 100 feet in depth are 

shown in Table 10-3 (Anaconda Collection-American Heritage Center). Anaconda, likely 

searching for shallow oxide feed for their Yerington mine, abandoned the claims sometime after 

1957. 

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Figure 10-1: Location of Historic Drill holes 

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Table 10-3: Anaconda Company 1955-1957 Drilling Highlights 


Feb-09 

Hole ID Total Depth (ft) Key Intercepts 

(Interval in feet and % Cu) 

Notes 

Mc 9 388 153-188: 0.52% Cu Bottomed in



not NI 43-101 compliant), described as an oxidized low-grade copper deposit which has been 

locally enriched by exotic copper (Heatwole, 1978). Anaconda’s resource calculations were 

developed into the mine plan supporting the 5.0 million tons at 0.30% Cu mined from the 

MacArthur pit by Arimetco during 1995-1997. A discussion of Anaconda’s drilling program 

with sampling protocol is presented in Appendix C. 

During 1987 to 1991, Pangea Explorations, Inc. located 304 unpatented lode claims and 

conducted an aggressive gold evaluation of the MacArthur area from the present day MacArthur 

pit westerly to the Gallagher area. Pangea’s program included over 549 rock chip samples, 

geologic and alteration mapping, followed by trenching two target areas (Adams, 1987). Eight 

trenches totaling over 1,420 feet were cut and sampled in the Gallagher area and four additional 

trenches totaling over 720 feet located in an undefined “north target.” Table 10-4 details some of 

Pangea’s exploration drilling results. Anomalous gold values (41 samples exceeding 0.015 Au 

oz/ton) led to a 15-hole / 2,110-foot reverse circulation drilling program with 1,310 feet in seven 

holes testing the Gallagher area. Pangea found the drilling results discouraging (best assay value 

of 0.026 Au oz/ton over 5 feet) and abandoned the property. 

Table 10-4: Pangea Exploration 1987-1991 Drilling Highlights 


Feb-09 

Interval 

Interval Length Gold Grade 

Hole ID (ft) (ft) (Au oz/ton) 

20-45 25 0.012 

MAC 91-1 165-175 10 0.013 

100-110 10 0.012 

MAC 91-2 130-145 15 0.016 

MAC 91-3 75-90 15 0.013 

45-55 10 0.011 

MAC 91-4 145-155 10 0.015 

MAC 91-5 90-100 10 0.011 

85-95 10 0.021 

100-110 10 0.014 

MAC 91-6 85-110 25 0.014 

5-15 10 0.015 

MAC 91-7 55-75 20 0.016 

MAC 91-8 105-115 10 0.016 

MAC 91-9 75-85 10 0.015 

MAC 91-10 60-80 20 0.014 

MAC 91-11 20-30 10 0.011 

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During the late 1980s through the late 1990s, Arimetco consolidated a major land position in the 

Yerington mining district consisting of over 8,500 acres including 85 patented claims. Arimetco 

entered the district to extract copper by heap leaching methods, with initial production from the 

Anaconda Yerington mine dumps, oxide stockpiles and Yerington mine vat leach tailings. 

Arimetco’s leach pads were located on the Yerington mine property approximately five miles 

south of the MacArthur property. During evaluation and mining of the MacArthur mine, 

Arimetco drilled an unknown number of holes as a check on Anaconda’s 1972 to 1973 drilling. 

Anaconda’s drilling and resource calculations provided the mine planning data for Arimetco’s 

MacArthur mine. Due to rising costs and depressed copper prices, Arimetco was forced to 

abandon their claim position and file for bankruptcy in 1999. 

In 2004, North Exploration located unpatented claims covering portions of the MacArthur 

property and the MacArthur pit that were leased to Quaterra in 2005. Subsequently, Quaterra has 

staked additional claims, bringing the current total to 470 unpatented lode claims over the project 

area. Quaterra’s current land position is displayed on Figure 4-2. 

10.3 HISTORIC MINING 

The MacArthur Project area has seen limited historic mining activity. The most recent activity 

occurred between 1995 and 1997, when Arimetco mined a limited tonnage (estimated 6.1 million 

tons) of surface oxide copper for heap leaching at the historic Yerington Mine site. No 

consistent, large-scale mining has occurred on the site. 

10.4 CURRENT DRILLING 

Although 2011 step out RC drilling at 500 foot centers continued to intersect acid soluble and 

primary copper mineralization, Quaterra focused mid-year on defining a resource that could 

support a positive mine plan. Drilling was therefore centered on an approximate one-half square 

mile area from North Ridge south to the MacArthur Pit, and the Gallagher area located west of 

the existing MacArthur Pit. Drill spacing was reduced to 250 foot centers on several drill fences. 

South-bearing angle holes tested the WNW, north dipping structural / mineralized grain and east- 

and west-bearing angle holes tested orthogonal structure resulting in the upgraded resource 

calculation reported Section 14.0 of this TR. 

During 2011, 81,651 feet of exploration drilling in 152 holes were completed including 69,890 

feet in 146 RC holes and 11,761 feet in six core holes. (See Figure 10-2) 

Also during 2011, 3,274.8 feet of PQ size core (3.35 inches) were drilled at 26 sites for 

metallurgical testwork. PQ holes twinned existing Quaterra RC and core holes. PQ holes were 

prefixed by “PQ-11” followed by the ID of the twinned hole. 

Quaterra’s complete exploration drill hole database, along with significant intercepts is listed in 

Appendix D. 

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10.5 SURVEYING DRILL HOLE COLLARS 

Drill hole locations are surveyed by Quaterra staff using a Trimble XHT unit, and are shown in 

Figure 10-2. 

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Figure 10-2: Drill hole Location Map 

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10.6 DOWNHOLE SURVEYS 

During 2011 five holes were downhole surveyed by International Directional Services, Elko, 

Nevada USA operating a surface recording Gyroscope. Downhole surveyed holes included: 

QM-163, 164, 165, 166, and 177. To date, downhole surveys have been completed on 57 of the 

Quaterra drill holes, and are now routinely done for holes greater than 1000 feet deep. 

10.7 CURRENT DRILLING METHODS AND DETAILS 

Quaterra has explored the MacArthur property with both reverse circulation (RC) and diamond 

core drilling methods. Reverse circulation holes have been drilled by Diversified Drilling LLC, 

Missoula, Montana, USA, DeLong Construction Inc., Winnemucca, Nevada, USA and by Leach 

Drilling Inc., Silver Springs, Nevada, USA. During 2007-2008 the core drilling was contracted to 

Kirkness Diamond Drilling of Dayton, Nevada, USA and Kirkness Brothers Diamond Drilling 

(aka KB Drilling Co, Inc.) of Carson City, Nevada, USA. Major Drilling America, Inc., Salt 

Lake City, Utah, conducted core drilling during 2009-2010. Core drilling during 2011 was 

contracted to Ruen Drilling Inc., Clark Fork, Idaho, USA. The RC crews ran one 10-12 hour 

shift per day; the core drill crews operated 24 hours per day. 

Quaterra has completed 204,656 feet of drilling in 401 holes since beginning drilling in 2007. 

Core holes total 40,233 feet in 58 holes and reverse circulation holes total 164,423 feet in 343 

holes. (Note that one previously listed, but abandoned 115 foot drill hole, has now been removed 

from the database and reported totals). Figure 10-3 show Quaterra's yearly exploration drilling 

footage by year. 

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Figure 10-3: Quaterra Exploration Drilling by Year 

During 2011, 69,890 feet in 146 RC holes and 11,761 feet in six core holes were drilled resulting 

in over 16,300 samples collected for analysis for total copper, acid soluble copper, ferric sulfate 

soluble (“QLT”), gold (selected samples), and trace elements (selected samples). Selected core 

was used to calculate rock quality designation (RQD) and measure bulk density. In addition 

3,274.8’ of PQ size core was drilled in 26 holes for metallurgical testwork. The total area 

covered by the MacArthur drilling is approximately 11,000 feet east-west by 6,000 feet northsouth 

at approximate drill spacing of 500 feet. Drill spacing reduces to approximately 250 feet 

within an approximate 1,500 feet east-west by 1,000 feet north-south within the northeast portion 

of the MacArthur pit and reduces to 250 foot spacing over portions of a 5000 foot square area 

north of the MacArthur pit. Historic Anaconda drilling spacing is 125 feet in the MacArthur pit. 

10.8 REVERSE CIRCULATION DRILLING SAMPLING METHOD 

All reverse circulation (RC) drilling is conducted with water added to eliminate dust. A 

percussion hammer with interchange sampling system has been used by the RC drill. Samples 

are collected in a conventional manner via a cyclone and standard wet splitter in 17-inch by 26inch 

cloth bags placed in five-gallon buckets to avoid spillage of material. Sample bags are premarked 

by Quaterra personnel at five-foot intervals and also include a numbered tag bearing the 

hole number and footage interval. Collected samples, weighing approximately 15 to 20 pounds 

each, are wire tied, and then loaded onto a ten-foot trailer with wood bed allowing initial 

draining and drying. Each day, Quaterra personnel, or the drillers at end of their shift, haul the 

sample trailer from drill site to Quaterra’s secure sample preparation warehouse in Yerington, 

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Nevada. Geologic logging samples are collected at the drill site in a mesh strainer, washed, and 

placed in standard plastic chip trays collected daily by Quaterra personnel. 

10.9 CORE DRILLING SAMPLING METHOD 

For 2011 exploration drilling core diameter was HQ (approximately 2.75-inch diameter. 

Following convention, at the drill site core was placed in wax-impregnated, ten-foot capacity 

cardboard boxes. Sample intervals vary from less than one foot to six feet, dependent upon rock 

consistency. Sample boxes were delivered to Quaterra’s secure sample warehouse in Yerington, 

Nevada by the drill crew following each 12-hour shift. 

PQ core drilling for metallurgical testwork followed similar protocol as exploration drilling. PQ 

core was placed in wax-impregnated, five-foot capacity cardboard boxes and delivered to 

Quaterra’s secure sample warehouse by the drill crew following each 12-hour shift. 

Special treatment was required for PQ core metallurgical samples to avoid undue oxidation prior 

to column testwork. PQ core was “quick-logged” without removing the core from core box or 

without breaking up the core. Prior to photographing, magnetic susceptibility and RQD 

measurements were collected. The entire core box was then sealed with plastic wrap to avoid 

oxidation. Shrink-wrapped core boxes were stacked on pallets, secured with plastic wrap and 

steel banding for shipment to METCON Research Laboratories in Tucson, Arizona USA. Time 

from core arrival at the core shed to plastic wrap of the core box was less than 24 hours. 

10.10 DRILLING, SAMPLING, AND RECOVERY FACTORS 

No factors were shown that could materially impact the accuracy and reliability of the above 

results. With few exceptions, core recovery exceeded 80% while RC recovery is estimated to be 

greater than 95%. 

10.11 SAMPLE QUALITY 

It is Tt’s opinion that Quaterra’s samples of the MacArthur Project are of high quality and are 

representative of the property. This statement applies to samples used for the determination of 

grades, lithologies, densities, and for planned metallurgical studies. 

It is the opinion of the author that during the period in 1972 to 1973 when Anaconda explored 

and drill tested the MacArthur property, the drill samples taken by Anaconda were representative 

of the deposit and the methodologies commonly used by the industry at that time. This statement 

applies to samples used for the determination of grades, lithology, and densities, as well as 

metallurgical performance, supported by similar determinations and conditions being carried out 

at that time at Anaconda’s Yerington mine operation and as referenced below in an internal 

Anaconda report (Heatwole, 1972), portions of which follow: 

“From March to November, 1972, over 225 holes were drilled... Approximately 33,000 feet of 

vertical hole and 13,000 feet of angle hole were drilled using percussion and rotary methods.” 

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The majority (62%) of the drilling, which was supervised by Anaconda’s Mining Research 

Department, was accomplished using Gardner-Denver PR123J percussion drills. The percussion 

drill was fitted with a sampling system designed by the Mining Research Department, which 

collected the entire sample discharged from the hole. The remainder of the drilling was done by 

Boyles Brothers Drilling Company using rotary and down-the-hole percussion equipment. The 

sampling system used by Boyles, especially during the early stages of drilling is not considered 

to be as accurate as the system designed by Mining Research. 

While no details are available regarding Anaconda’s exact assaying protocol and quality control 

during drilling at the MacArthur property, an interview conducted by Quaterra personnel in 

October 2008 with Mr. Henry Koehler, Anaconda’s Chief Chemist during the 1960s and 1970s, 

confirmed that the techniques and procedures implemented conformed to industry standards for 

that era. Mr. Koehler was employed in Anaconda’s analytical laboratory from 1952 to mine 

closure in 1978. He currently resides in Yerington, Nevada. 

Figure 10-4: Letter from Mr. Henry Koehler 

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11 SAMPLE PREPARATION, ANALYSES AND SECURITY 

Tt has reviewed all of the Quaterra sample preparation, handling, analyses, and security 

procedures. It is Tt’s opinion that the current practices meet NI 43-101 and CIM defined 

requirements. Following a Tt recommendation, standards are stored in a locked and secured area 

11.1 RC SAMPLE PREPARATION AND SECURITY 

RC sample bags, having been transported on a ten-foot trailer by Quaterra personnel from the 

drill site to the secure sample warehouse, are unloaded onto suspended wire mesh frames for 

further drying. Diesel-charged space heaters assist in drying during winter months. Once dry, 

sets of three samples are combined in a 24- by 36-inch woven polypropylene transport (“rice”) 

bag, wire tied, and carefully loaded on plastic lined pallets. Each pallet, holding approximately 

13 to 15 rice bags, is shrink-wrapped and further secured with wire bands. Quaterra’s samples 

were shipped via UPS Freight to Skyline Assayers & Laboratories (Skyline), Tucson, Arizona 

USA through 2008. During the 2009-2010 drill campaign, Skyline dispatched a transport truck 

from Tucson to collect samples. In 2011, Skyline established a sample preparation facility in 

Battle Mountain, Nevada, from which trucks were dispatched to pick up Quaterra’s drill samples 

under a chain of custody protocol. Following sample preparation in the Battle Mountain facility, 

Skyline ships a representative pulp sample to the Skyline laboratory in Tucson, Arizona for 

analysis. 

Complying with earlier recommendations from Tt, Quaterra now weighs each shrink-wrapped 

pallet of samples prior to departure from Yerington. Rejects and pulps are returned to Quaterra 

and stored under cover in a secure location. 

11.2 CORE SAMPLE PREPARATION AND SECURITY 

Drill core, having been transported at end of each shift by the drill crew to Quaterra’s secure 

sample warehouse, is logged by a Quaterra geologist who marks appropriate sample intervals 

(one to nominal five feet) with colored flagging tape. Each core box, bearing a label tag showing 

drill hole number, box number, and box footage interval, is then photographed. Rock quality 

designations (RQD), magnetic susceptibility, and recovery measurements are taken. Core 

preceding drill hole QMCC-20 was sawed in half by Quaterra personnel; core holes QM-026, 

QM-036, QM-041, QM-046, and QM-049 were split in half using a hydraulic powered blade at 

the warehouse by Quaterra personnel (for approximately six months through core hole QMCC- 

20, core was sawed rather than hydraulically split). One half of the split was bagged in 11- by 

17-inch cloth bags for assay while the other half was returned to the appropriate core box for 

storage in the sample warehouse. Approximately five to six cloth sample bags are combined in a 

larger 24- by 36-inch transport polypropylene (“rice”) bag, wire tied, and carefully loaded on 

plastic lined pallets. Each pallet, holding approximately 13 to 15 rice bags, was shrink-wrapped 

and further secured with wire bands for shipment to Skyline in Tucson. The same chain of 

custody protocol is used for both RC and core samples. 

Following geologic logging and RQD measurements, the core portions of holes QM-99, QM- 

100, and QM-109 (2009-2010) and QM-163, QM-164, QM-165, QM-177 and QM-185 (2011 

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program) were strapped and shrink wrapped on pallets for shipment to ALS Minerals laboratory 

in Reno, Nevada. Core samples were picked up from the warehouse by a Reno, Nevada-based 

ALS Minerals driver, and sample pallets were weighed upon receipt by the laboratory. ALS 

personnel sawed the core in half, one half for assay at the ALS laboratory, storing the other half 

in the core box for return to Quaterra. Chain of custody procedures for ALS Minerals follow the 

format described for Skyline. 

Following geologic logging, magnetic susceptibility and RQD measurements, and photography, 

PQ core for metallurgical testing was shrink-wrapped in its cardboard core box, stacked on 

pallets, shrink-wrapped together, wire banded, and weighed. Pallets were shipped to METCON 

Research Laboratories, Tucson, Arizona via UPS Ground. Chain of Custody was signed upon 

departure from Yerington and receipt in Tucson. 

11.3 SAMPLE ANALYSIS 

During 2007, 12 drill holes (core) were analyzed at American Assay Laboratories (AAL) in 

Sparks, Nevada, USA. AAL is ISO/UEC 17025 certified as well as a Certificate of Laboratory 

Proficiency PTP-MAL from the Standards Council of Canada. 

With sample submission-to-reporting time exceeding two months at AAL, Quaterra elected to 

use Skyline Assayers & Laboratories (Skyline) and ISO certified assay lab in Tucson, Arizona, 

USA for all further analytical work. Samples submitted to AAL were re-assayed (pulps or 

rejects) by Skyline for consistency of the data set. 

Core from drill holes QM-99, QM-100, and QM-109 (2009-2010) and QM-163, QM-164, QM- 

165, QM-166, QM-177 and QM-185 (2011 program) were submitted to ALS Minerals, Sparks, 

Nevada, USA. ALS Minerals is an ISO registered and accredited laboratory in North America. 

Quaterra samples arrive at Skyline via UPS truck freight and in 2009-2010 by a transport truck 

dispatched from Tucson by Skyline. A Quality Assurance and Quality Control Assay Protocol 

have been implemented by Quaterra where one blank and one standard are inserted with every 

18 drill hole samples going into the assay stream. The Skyline assay procedures are as follows: 

• For Total Copper: a 0.2000 to 0.2300 gram (g) sample is weighed into a 200-milliliter 

(ml) flask in batches of 20 samples plus two checks (duplicates) and two standards per 

rack. A three-acid mix, 14.5 ml total is added and heated to about 250°C for digestion. 

The sample is made to volume and read on an ICP/AAS using standards and blanks for 

calibration. 

• For Acid Soluble Copper: a 1.00 to 1.05 g sample is weighed into a 200 ml flask in 

batches of 20 samples plus two checks (duplicates) and two standards per rack. Sulfuric 

acid (2.174 l) in water and sodium sulfite in water are mixed and added to the flask and 

allowed to leach for an hour. The sample is made to volume and read on an ICP/AAS 

using standards and blanks for calibration. 

• For Ferric Soluble Copper (QLT): uses an assay pulp sample contacted with a strong 

sulfuric acid-ferric sulfate solution. The sample is shaken with the solution for 30 minutes 

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at 75ºC, and then filtered. The filtrate is cooled, made up to a standard volume, and the 

copper determined by AA with appropriate standards and blanks for calibration. 

• For Sequential Copper Leach: consists of four analyses: Total Copper, Acid Soluble 

Copper, Cyanide Soluble Copper, and the difference, or Residual. Following analysis for 

Total Copper and Acid Soluble Copper, the residue from the acid soluble test is leached 

(shake test) in a sodium cyanide solution to determine percent cyanide soluble minerals. 

The Sequential Copper Leach is a different approach to the Ferric Soluble Copper (QLT) 

leach, with possible greater leaching of certain sulfides (e.g. chalcocite or bornite) during 

the cyanide leach step. 

Beginning in 2009, Quaterra requested 34-element trace element geochemistry from Skyline on 

selected samples which were analyzed by ICP.OES Aqua Regia Leach. 

During 2009-2010 Quaterra core samples were picked up at Quaterra’s warehouse facility by 

ALS Minerals personnel and transported to ALS Minerals laboratory in Sparks, Nevada, USA. 

ALS Minerals personnel sawed the core, saving one-half for return to Quaterra. ALS assayed 

core for trace element geochemistry with 48-element Four Acid “Near-Total” Digestion. 

In keeping with Tt recommendations, beginning in 2009, Quaterra began a program to re-assay 

selected samples when blanks, standards, or repeat assays exceeded or were below the expected 

values by 15%, or blanks returned an assay of >.015% Cu. The QC program now re-assays 

standards outside +/- 2 standard deviations of the expected value, repeat assays +/- 15% of the 

original assay, and blanks greater than .015% Cu. 

11.4 LEACH ASSAY ANALYSIS 

Both Sequential copper leach assays and QLT leach assays, when combined with column leach 

tests can be indicative of actual heap leach recoveries. Historically, sequential copper leach 

assays were not performed on samples at MacArthur. Section 6.4 discusses the problems 

encountered by previous operators while leaching ore material from the MacArthur pit. Since 

previous operators were unable to explain the longer leach times and low solution head grades 

they encountered, Tt recommended that Quaterra perform sequential copper leach assays on 

some of the available sample coarse rejects. While only early results were available for the 2009 

TR, Table 11-1 shows a January 2011 updated summary of the total copper, acid-soluble copper 

(ACu), and cyanide-soluble copper (CNCu) quantities categorized by mineralized zones. The 

acid-soluble fraction of total copper is greatest in the oxide zone. The cyanide-soluble fraction of 

total copper is greatest in the chalcocite/oxide zone where the dominant species of copper 

mineral is chalcocite. In the primary sulfide zone, both acid- and cyanide-soluble fractions of 

total copper are low due to high levels of chalcopyrite. 

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Table 11-1: Sequential Copper Leach Assay Results 

QUATERRA ALASKA, INC. – MACARTHUR PROJECT 

January 2011 

Average Values 

Mineralized Zone %Cu %ACu %CNCu ACu : Cu CNCu : Cu 

% Soluble 

Cu 

# of 

Samples 

Oxide 0.185 0.103 0.015 0.56 0.08 64% 213 

Chalcocite/Oxide 0.252 0.065 0.084 0.26 0.33 59% 281 

Primary 0.186 0.019 0.030 0.10 0.16 27% 60 

Tt proposed that Quaterra performed either, standard CU assays, warm H2SO4 assay, and QLT 

or standard sequential copper leach assays on all drill hole samples that exceed 0.10% Cu for all 

future drilling programs. This data will help Quaterra to better understand potential 

mineralogical differences between the oxide, secondary, and primary mineral zones as well as 

help link column leach test composites with in situ material to better predict heap leach 

performance. 

Beginning with drill hole QM-086 in December 2009, Quaterra continued to request analyses for 

total copper (Cu or TCu) from all drill samples. Analyses for acid soluble copper (ACu) and for 

ferric sulfate leach aka Quick Leach Tests (QLT) were requested for drill samples (plus an 

additional 50 feet downhole) containing visible green or black copper or containing chalcocite. 

These analyses were completed by Skyline Laboratories, Tucson, Arizona for all reverse 

circulation drilling. A high acid soluble copper to total copper ratio indicates that leachable oxide 

copper is present. QLT minus acid soluble offers an estimate of acid soluble (leachable) sulfide 

copper, i.e. chalcocite. The Cu-ACu-QLT analysis combination is an alternative approach, rather 

than using the Sequential Leach analysis (Cu, ACu, cyanide soluble copper, then calculate 

residual). 

Skyline Laboratories performed QLT analyses from both core and reverse circulation drill 

samples at the onset of Quaterra’s exploration program in 2007. In order to complete the QLT 

analyses through drill hole QM-085, Skyline analyzed an additional 2,747 pulps representing 

both core and reverse circulation drill footages. 

Table 11-2 summarizes the results of the Cu, ACU, and QLT testing for those intervals 

containing all three assays and where the Cu value is >0.1% Cu. It should be noted that no ACU 

or QLT assays were included in the data received for the Anaconda drilling within the 

MacArthur oxide deposit. The data shown in Table 11-2 therefore reflects only the Quaterra 

drilling and reflects averages over the entire project area. 

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Table 11-2: Ferric Sulfate Leach (QLT) Assay Results 

QUATERRA ALASKA, INC. - MACARTHUR PROJECT 

January 2012 

Averages for samples with Cu>.1%Cu with ACU and QLTCU Assays 

Mineralized Zone %Cu ACu QLTCu ACu : Cu QLTCu:Cu % Soluble Cu 

# of 

Samples 

Oxide 0.236 0.116 0.128 0.425 0.471 47% 1585 

Chalcocite/Oxide 0.341 0.066 0.143 0.215 0.401 40% 2576 

Primary 0.315 0.029 0.063 0.096 0.171 17% 486 

Further metallurgical work is expected to provide a better understanding of the differences noted 

when comparing the QLT and sequential leach method results. 

11.5 QUALITY CONTROL 

As part of the quality control program, 675 standards and 622 blanks were submitted (Table 

11-3) along with 15,063 individual drill hole samples to Skyline Laboratories. Additionally, 87 

standards and 85 blanks were submitted along with 1,748 core samples to ALS Mineral Labs in 

Reno. 

Lot failure criteria were established as any standard assaying beyond two standard deviations of 

the expected value, or any blank assay greater than 0.015% Cu. Failed lots were reviewed and lot 

samples were selected for reassay. Results indicated that all original assays, with the exception of 

4, in which sample numbers had been switched, would be accepted as originally received. 

Table 11-3: MacArthur 2011 QA/QC Program Results 

Skyline Labs ALS Mineral Labs 

Total Drill Hole Samples 15,063 1,748 

Submitted Standards 675 87 

Failed Standards 27 3 

% Standards Failure 4.0% 3.4% 

Submitted Blanks 622 85 

Failed Blanks 7 0 

% Blank Failure 1.1% 0% 

Check assays from ALS Mineral Labs compared well with Skyline assays, providing additional 

confidence in the assay database, as shown in Figure 11-1. 

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Figure 11-1: MacArthur Check Assay Results 

11.6 REVIEW OF ADEQUACY OF SAMPLE PREPARATION, ANALYSES, AND SECURITY 

During the visit to the project in 2011, Dr. Bryan observed geologic logging and data entry of 

drill data following an established protocol (Figure 11-2), and procedures for manually creating 

geologic sections from the drill data (Figure 11-3). 

Figure 11-2: Reviewing Established Protocol for Data Entry 

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Figure 11-3: Manually Creating Geologic Sections from the Drill Data 

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12 DATA VERIFICATION 

Dr. Rex Bryan of Tt conducted a site visit to the MacArthur Project area and Quaterra’s field 

office in Yerington, Nevada on September 9, 2011. During this visit Quaterra staff discussed the 

history of the project, presented all requested data, answered questions posed by Tt, presented 

the current geologic interpretation of the MacArthur deposit, and guided Dr. Bryan on a field 

examination through the MacArthur property which included observing drill sample collection 

during reverse circulation drilling. This section details the results of Tt’s verification of existing 

data for the MacArthur Project. 

12.1 HISTORIC DATA CHECK 

Tt did not collect independent samples to corroborate historic data. It is Tt’s opinion that the 

previous owners of the MacArthur Project area were competent established companies that 

followed industry standard practices for drilling, sampling, and assaying according to the 

industry standards in place at the time of the work. However, Quaterra has completed 

verification work on the historic data by re-assaying, when material was available, and twin hole 

drilling. 

As an assay check on the historic Anaconda drilling within the confines of the current 

MacArthur pit, Quaterra twinned nineteen Anaconda holes using both reverse circulation and 

core drilling methods (Table 12-1). The attached histogram (Figure 12-1) contains information 

on 57 total holes: 38 Quaterra and 19 Anaconda. It provides a comparison of average copper 

grades between the 1972-1973 Anaconda drilling (all as dry drilling, capturing 100% of the dry 

sample) and Quaterra’s twin holes (wet sample recovery for all Quaterra reverse circulation 

drilling). Some of the twin holes drilled by Quaterra are angled whereas the corresponding 

Anaconda hole was drilled vertically. For these twin angle-drilled holes, the intercept displayed 

in Figure 12-1 is the length-weighted average over the projected vertical interval. The 

abbreviations Q-aRC and Q-bRC are first and second twins of existing holes. 

A complete discussion of the twin hole program is available in the "MacArthur Copper Project 

NI 43-101 Technical Report", dated Feb 17, 2009. 

12.2 CURRENT DATA CHECK 

Tt has made several data checks and verifications of Quaterra work that has been performed for 

the MacArthur Project. These checks include validation of assays from Skyline and comparing 

geologic field logs with drill hole data. No discrepancies have been found. 

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Table 12-1: List of Twin Holes Drilled By Quaterra 

QUATERRA ALASKA, INC. – MACARTHUR PROJECT 

Twin Group Anaconda Hole 

February 2009 

Quaterra 

Twin Core 

Hole 

1 M120-C50-1 QMT-4 

Quaterra Twin 

aRC Hole 

2 M120-C50-2 QMT-5 QMT-5aR 

3 M165-K-1 QMT-11 QMT-11aR 

4 M172.5-I-1 QMT-8 QMT-8aR 

5 M195-M-1 QMT-13 QMT-13aR 

Quaterra Twin 

bRC Hole 

6 M195-M-2 QMT-14 QMT-14aR QMT-14bR 

7 M205-G-2 QMT-6 

8 M210-K-1 QMT-10 QMT-10aR QMT-10bR 

9 M210-O-1 QMT-15 QMT-15aR 

10 M270-Q-1 QMT-17 QMT-17aR QMT-17bR 

11 M270-S-1 QMT-18 QMT-18aR QMT-18bR 

12 M30-K-1 QMT-12 QMT-12aR 

13 M45-C1-1 QMT-1 QMT-1aR QMT-1bR 

14 M45-C1-2 QMT-2 QMT-2aR 

15 M75-I-1 QMT-9 

16 M90-B-1-2 QMT-3 QMT-3aR 

17 M-90-G-4 QMT-19 

18 M90-O-1 QMT-16 QMT-16aR 

19 M95-G-1 QMT-7 

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12.2.1 Adequacy of Data 

Figure 12-1: Twin Hole Charted Results 

It is Tetra Tech’s opinion that the data collection of both historic and modern data by Quaterra is 

adequate for the use of a 43-101 resource for the following reasons: 

• The sampling is representative of the deposit in both survey and geological context 

• The drill hole cores have been archived and are available for further checking 

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13 MINERAL PROCESSING AND METALLURGICAL TESTING 

The MacArthur deposit generally consists of an oxidized copper capping transitioning through a 

mixed oxide/secondary copper interface into primary sulfides at depth. Essentially all 

metallurgical testwork to date has been conducted on the copper oxide resources with a few tests 

having been performed on mixed oxide/sulfide material. 

The MacArthur Project has a long history of metallurgical bottle roll and column testwork from 

1976 through 2011. Historical test work by Anaconda in 1976 included bottle roll and column 

leach tests on samples collected from surface trenches. Arimetco performed a number of bottle 

and column leach tests on surface samples between 1992 and 1995 using several different 

metallurgical laboratories. Quaterra performed bottle roll and column tests between 2010 and 

2011 through METCON Research in Tucson, Arizona. A list summarizing this testwork can be 

found in Table 13-1 at the end of this section. 

Of significance, Anaconda operated a vat leach facility processing oxide ore from the Yerington 

Pit, the results from which were documented over the many years of operation. Arimetco also 

operated a number of leach pads between 1989 and 1995 treating oxide and transition ores mined 

from the Yerington Pit. However, between 1994 and 1997, approximately 6.1 million tons of ore 

was mined from the MacArthur Pit and hauled Run of Mine (ROM) to the Arimetco pads for 

processing. This commercial operational database for both the vat and heap leach operations was 

significant since both Yerington and MacArthur ore deposits are very similar in origin, geology 

and mineralization. A summary of several years of data from the vat leach operation is available 

for review. 

A review of the METCON metallurgical test work shows good copper extraction but variable 

acid consumption spatially throughout the deposit. METCON column test work (32 columns) 

conducted in 2011 using material from 32 different PQ core drill holes (rather than material 

taken during surface sampling) was completed for the PEA. The drill holes provided reasonable 

spatial representivity of the MacArthur resources in all the deposit area. The METCON column 

study completed in 2011 is available for review. 

Combined, the 2011 METCON study, the Anaconda vat leaching data, and the Arimetco 

commercial leach pad data provided sufficient metallurgical information to gain a preliminary 

level of confidence used in developing this PEA Study. However, to achieve the level of 

confidence for a prefeasibility study (PFS), additional metallurgical test work is necessary to 

better understand acidification techniques and the resultant copper extraction spatially in 

mineralized resource contained within the mine plan. This metallurgical test work will be 

undertaken during the PFS. Recommendations for this test program are provided in Section 26 of 

this report. 

The following sections discuss the criteria upon which generation of the heap leach design 

parameters for the PEA were made. The design parameters are summarized in sub-section 13.9 

Heap Leach Design Criteria. 

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13.1 OXIDE ORE COPPER EXTRACTION 

Predicted copper extraction and acid consumption was derived from the existing metallurgical 

data base, METCON columns and Arimetco historical information. Figure 13-1 below shows 

column copper extraction versus grade during a 120 day leach cycle. The 32 METCON columns 

average 60% extraction, which is globally near the extraction achieved by Arimetco at a similar 

copper grade. Half of the columns averaged 0.11% copper head grade which likely provides a 

downward bias on copper extraction. As grade increases, copper extraction increases. Assuming 

a 0.15% copper cutoff grade, Figure 13-1 shows column copper extraction of 65%. Using a 

permanent heap leach pad, extraction is predicted to increase during residual leaching of the 

overlaid pads, greater than offsetting solution copper inventory buildup in the pad. 

Figure 13-1: Comparison of Grade versus Copper Recovery Oxide Leach Ore 

The 32 METCON column tests showed that the leaching performance in the old MacArthur Pit 

area provided higher copper extraction and lower acid consumption compared to both the 

Gallagher Pit area and the North and Northwest MacArthur Pit areas. The variation in leach 

performance is not fully understood and will be addressed with future drilling and metallurgical 

test work in the PFS (discussed in Section 26). 

The nine columns from the historic MacArthur Pit area averaged 83.9% extraction. Based on the 

65% average column extraction of the 32 columns using a 0.15% cutoff grade, MacArthur Pit 

resources were conservatively predicted to achieve 70% copper extraction while all other pit 

areas were predicted to achieve 65% extraction. Acid consumption for the MacArthur Pit 

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material only was also reduced from the estimated global acid consumption of 35 pounds of acid 

per ton of oxide ore to 30 pounds per ton. Oxide ore from the other pit areas was projected at 35 

pounds of acid per ton of ore. Mixed oxide/secondary sulfide ores were projected to consume 30 

pounds of acid per ton of ore. 

The project life of mine (LOM) mine plan shows that, of the 271 million tons of ore, 132 million 

consists of the MacArthur Pit oxide ore (49%). The plan also shows that 68.4% of the total ore is 

oxide mineralization, mixed oxide and secondary sulfide making up the remaining 31.6 percent. 

Of importance to project economics are the timing and sequence of ore production. MacArthur 

Pit oxide ore constitutes 90 percent of the material leached during the first 7 years of mine life. 

MacArthur oxide ore having the highest copper leach extraction, the lowest acid consumption, 

and the lowest strip ratio in these first years’ works to optimize project economics. 

13.2 OXIDE ORE ACID CONSUMPTION 

Column test work assumed the use of an acid cure application followed by continued 

acidification during leaching/rinsing of the columns. During the cure stage, 31.59 pounds of acid 

per ton of ore was added. Following the acid cure, leaching of most columns consumed almost 

an equal amount of additional acid during the 120 day leach cycle. Most columns were operated 

between 1.5 and 1.6 pH during this leach cycle. It is probable that all 32 columns were over 

acidified both during the acid cure and leaching which resulted in excess acid consumption, 

averaging 57.3 pounds of sulfuric acid per ton of ore processed. 

Using leach test results from only the 16 columns at a cut-off grade of 0.15% copper, and 

disregarding the test work results from the Gallagher Pit zone, acid consumption was determined 

to be 45.4 pounds of sulfuric acid per ton of ore. These test work results, taken in conjunction 

with qualified opinion predicts that acid consumption may be reduced 20% to 36.3 pounds of 

acid per ton of ore considering the column over acidification that was realized combined with 

shortening of the leach cycle time to 90 days. Arimetco added 25 to 30 pounds of acid per ton of 

ore with 7.7 pounds of acid consumed per pound of copper produced. However, since acid 

consumption appears to increase on the periphery of the MacArthur Pit, average acid 

consumption for all ore was estimated at 35 pounds of acid per ton of ore. 

13.3 TRANSITION ORE EXTRACTION AND ACID CONSUMPTION 

Research for prediction of copper leach extraction from secondary sulfides (chalcocite) in the 

transition ores is limited to one METCON column. A number of bottle roll tests with high levels 

of secondary copper were also run but bottle roll tests are considered index tests and do not 

produce data with an acceptable level of confidence on their own for heap leach design purposes. 

The total grade of the METCON column #4 was 0.363% copper with a cyanide soluble copper of 

0.203% (secondary sulfide). Leach extraction of the secondary copper values was 56%, the 

extraction kinetics being slower than the oxide columns which is typical of secondary sulfide 

leaching. Leach extraction after 120 days was still significant and would continue in practice 

through the residual leaching of lifts as this material is overlaid by fresh ore. 

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The total head iron content was 3.87%Fe with a tail residue of 3.32%Fe, showing an iron leach 

extraction of 6.22%. Test results from this column showed the least continuing acid consumption 

and iron extraction. Acid added during the cure was 32.5 pounds of acid per ton of ore. A total of 

45.56 pounds of acid per ton of ore were consumed during this 120 day column test. The pH of 

the leach solution on day one of the column test was 0.43 indicating that the column was likely 

over acidified. 

During the leach cycle the column pH ran between 1.45 and 1.55 and was much easier to 

maintain at this level. 

Ferric iron concentration was 14.3 g/l the first day of rinsing which supplies ferric iron for 

chalcocite leaching. The ferrous iron was near zero after about 20 days of leaching showing that 

first stage chalcocite leaching was complete. The solution oxidation/reduction potential (ORP) 

remained about 650 mV after 20 days, ideal for second stage chalcocite leaching. 

The head screen analysis of the one secondary sulfide column tested was coarser than the 

materials in the other 31 columns. This column also showed minimal chemical degradation. The 

head screen analysis was significantly coarser than the column averages and very little chemical 

degradation occurred. Chalcocite may tend to be more disseminated within the host rock than 

oxide ore. Although the copper grade in the column is not high, some acid will be generated 

during residual leaching as the second stage of chalcocite (covellite) is slowly leached resulting 

in elemental sulfur formation. Therefore, considering a shorter leach cycle time, acid 

consumption for secondary sulfide ore leaching was predicted to be 30 pounds of sulfuric acid 

per ton of ore. Copper leach extraction with residual leaching is predicted at 60 percent. 

13.4 LEACH CYCLE TIME 

A review of the column test work shows that copper extraction was nearing completion after 75 

to 90 days. Overall extraction kinetic curves of the 32 columns were marginally slower than 

typical oxide leach columns, perhaps due to some oxide mineral dissemination within the host 

rock. A 90 day leach cycle was selected. This cycle time was also chosen to minimize the 

continuing acid consumption with little copper extraction occurring over the remaining 30 days 

of the 120 day leach as realized in the columns. Residual leaching beneath an overlaid lift will 

not see the 6 to 8 g/l acid concentration of a new lift. Once overlaid, the lift will see between 2 

and 4 grams per liter, still promoting leaching but at significantly lower acid consumption, 

thereby optimizing overall copper extraction and acid consumption globally. 

13.5 LEACH SOLUTION APPLICATION RATE 

Considering the chemical degradation experienced during the column leach tests and with the 

Arimetco leach pads, a leach application rate of 0.0035 gpm per square foot was selected. With 

multiple lift overlays, the leach pad will see continuing consolidation by chemical degradation 

and ore column weight. Thus permeability will tend to decrease probably limiting the leach 

application rate to 0.0035 gallons per minute per square foot. The new leach pads can still be 

leached at higher application rates for the first 20 to 30 days to optimize copper extraction 

kinetics while the bulk of the pads can be operated at a lower application rate. 

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13.6 PAD HEIGHT 

Production lift height was selected to be 20 feet even though a 15 foot lift may be preferable 

metallurgically. The pad footprint, leach flow rate, and capital and operating costs increase 

proportionally with reduced lift heights. Extraction performance at the grade of this ore with a 15 

foot high lift height would likely not justify these significant capital and operating costs. 

Additional test work in the pre-feasibility stage will supply better data to refine the lift height. 

13.7 PLS FLOW RATE AND PLS GRADE 

Assuming no intermediate leach solution recycle, the PLS flow rate is estimated at 10,400 

gallons per minute with a PLS grade of 1.0 g/l copper, including raffinate recycle of 0.1 g/l 

copper considering 90% SX recovery of copper. M3 Engineering confirmed the flowrate of 

10,400 gpm and PLS grade of 1 g/liter based on leach duration, irrigation rates and lift height. 

The PLS grade was determined by copper recovery and PLS flow rate. 

13.8 PARTICLE SIZE TO HEAP LEACH 

Historic test work provides limited ROM data for copper extraction and acid consumption. 

However, MacArthur ROM ore was successfully processed by Arimetco with good copper 

extraction and acid consumption, supporting the ROM leaching approach. The proposed Phase 

II PFS metallurgical program will address particle size vs. copper extraction and acid 

consumption (refer to section 26 for additional detail). 

13.9 HEAP LEACH DESIGN CRITERIA 

As a result of studying and analyzing all metallurgical test work to date, the following design 

criteria were developed for the MacArthur Project PEA. 

• Annual leach ore mining rate -15,000,000 tons 

• Daily leach ore mining rate - 41,095 tons 

• ROM truck dumping direct to the leach pad 

• Acidification procedures to be determined during the pre-feasibility test work 

• Leach pad slope at 1.75 to 1 with step backs between each lift 

• One inner-lift liner at mid-point of the final pad elevation 

• Leach cycle time-90 days 

• Leach solution application rate-0.0035 gallon per minute per square foot 

• Lift height-20 feet 

• Individual leach module lift size- 250 feet by 600 feet by 20 feet deep 

• PLS grade-1.0 grams per liter 

• PLS flow rate- 10,400 gallons per minute 

• Ore bulk density- 125 pounds per cubic foot 

• MacArthur Pit oxide ore 

o Copper extraction- 70% 

o Acid consumption 30 pounds of acid per ton of ore 

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• Other oxide ore 


o Acid consumption- 35 pounds of acid per ton of ore 

• Transition sulfide ore 


o Acid consumption-30 pounds of acid per ton of ore 

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Date Laboratory Prepared for Sample type 

1995 Leach Inc., Tucson,AZ Arimetco, Inc. 


Head 

Grade 

% Tcu 

Bulk 2,450 lb 

from 

MacArthur 0.302 

Bulk 2,450 lb 

from 


Bulk 2,450 lb 

from 


Table 13-1: MacArthur Historical Test Work 

Head 

Grade 

% AsCu Treatment 

% 

Recovery 

Tcu 

% 

Recovery 

AsCu 

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Cumulative 

PLS Grade 

gpl 

Cure Acid, 

lb/ton ore 

Acid 

Consumption 

lb. / ton ore 

Acid 

Consumption 

lb acid/lb Cu 

Extracted Irrigation Rate gpm/ft.2 Remarks 

BM-1 Crush-6" 24"x 8.2' 

c olumn 65.5% 0.77 20.5 34.0 8.6 Initial 0.0045 gpm/ft.2 Cu present as Chrysocolla. 

BM-2 Crush-3" 12"x9.3' 

c olumn 75.0% 0.52 24.5 35.5 6.8 for 42 days , then reduced No malachite, azutite or Cuprite 


BM-3 Crush-1"; 8"x 9.9' 

c olumn 84.6% 1.32 36.2 59.3 10.1 to 0.0030 gpm/ft2 for 18 days All samples - acid cure 7 days 

Acid Consumption 8 hour test of 

1995 55 

sample, with pH of 1.8 

Calculated Head Grades Acid Consumption Calculate 

Acid consumption after SX/EW Credit 

1992 

1992 

1990 

1990 

McClelland, Sparks, NV-1 

Summary Progress Report 

McClelland, Sparks, NV-2 

Summary Progress Report 

Mountain States, Tucson, 

AZ - Final Report 

Mountain States, Tucson, 

AZ - Final Report 

1989 Bateman Metallurgical 








Mine Development 

Associates Bulk 2,300 lbs 0.8315 0.6285 

Mine Development 

Associates Bulk 2,180 lbs 0.8315 

Total Cu 

0.6285 

Crush-6",24"x10' column,no 

acid agglom 19.5% 25.9% N/A 0 17.5 7.6 N/A 

Crush-6",24'x10' column, 

acid agglom 54.2% 71.7% N/A 30 33.4 5.5 N/A 

MacArthur Mining & 

Processing Bulk 520lb BM- 

Company 

14 0.56 Crush-2 1/2' 10"x10' Column 81.6% N/A 30 30.2 3.3 0.004 gpm/ft2 

MacArthur Mining & 

Processing Bulk 520lb BM- 

Company 

15 0.51 Crush-2 1/2' 10"x10' Column 75.6% N/A 30 27.7 3.6 0.004 gpm/ft3 

Calculated Head Grade Recovery after 48 days 

Timberline 

Minerals, Inc. 

Timberline 


Timberline 


Timberline 


Timberline 


1992 Arimet c o Arimet c o 

1992 Arimet c o Arimet c o 

1976 Anaconda 

1976 Anaconda 

Phase 1 

Trench-8 tons 

TMI-A 1.114 0.862 

Phase 1 

TMI-B 

Phase 2 

Trench-8 tons 

0.334 0.269 

TMI-A 22 lbs 1.108 0.862 

Phase 2 TMI- 

B 22 lbs 0.345 0.269 

Phase 3 

TMI-B 22 lbs 0.338 0.269 

Phase 1 Bottle Roll, Crush - 

2" (average of 2 runs) 79.2% 84.1% N/A 175 85.9 4.9 


2" (average of 2 runs) 38.0% 51.1% N/A 144 49.8 19.6 


2" (average of 4 runs) 67.9% N/A 70 & 105 81.75 5.4 


2" (average of 3 runs) 48.2% N/A 57.5 35.8 10.8 


2" (Run #6) 40.6% 51.1% N/A 57.5 36.1 13.2 

Bulk Sample A 

(TMI-A) 2,100 

Crushed to minus 3 inches, 

MacArthur Mining & lbs duplicate 

column leach 18 inch dia by 

Processing samples 

Sample B (T MI- 

0.675 0.485 

15 ft 44.2% 61.6% N/A 

B) 3,500 lbs 

Crushed to minus 3 inches, 

MacArthur Mining & duplicate 

column leach 24 inch dia by 

Processing samples 0.335 0.26 

15 ft 22.8% 29.3% N/A 

MacArthur Mining & sample B with 

Processing chloride ion 0.216 0.167 49.3% 64.0% N/A 57.5 38.1 17.9 

Calculated Head Grades 

Samples from 

existing 

trenches N/A 

Samples from 

existing 

trenches N/A 

Composit e 

Samples by 

ore type by 

Anaconda 

MRD 76-118-A 

Composit e 

Samples by 

ore type by 

Anaconda 

0.48 0.4 

MRD 76-118-B 0.85 0.74 

50 & 100 gpl 

acid 

preconditioning 32.85 5.5 .004 gpm/ft2 

51 & 100 gpl 

acid 

preconditioning 10.95 7.2 .004 gpm/ft3 

ROM? 12'"x10' column A, 

Leached 93 days 82.7% N/A N/A 7.8 

ROM? 12'"x10' column B, 

leached 58 days 91.6% N/A N/A 8.1 

Bottle Role Crushed to -3/8 

in., 2 kg samples leached 

for 120 hours 83.24% N/A N/A 80.7 12.5 



for 120 hours 87.0% N/A N/A 86.0 9.4 

Composit e 

Samples by 

ore type by 


Anaconda 


1976 Anaconda 

Duplicate of above tests 

MRD 76-118-C 0.14 0.08 

for 120 hours 63.3% N/A N/A 72.0 38.8 

1976 done in Tucson 


MRD 76-118-A 1 0.88 89.1% N/A N/A 134.0 6.4 

1976 done in Tucson 


MRD 76-118-B 2.17 2.16 96.5% N/A N/A 114.0 2.6 

1976 done in Tucson MRD 76-118-C 0.23 0.15 65.3% N/A N/A 64.0 14.1 

NOTE:Acid consumption in 

tests give as Lbs/Ton ore. 

Lb acid/Lb Cu calculated as 

((head gradeX20)X% 

recovery)/Lbs Acid/Ton 

Calculated Values 

One bulk sample split to 2 tests. 

Tests were in progress 51 days. 

Tests were in progress and not 

completed. Recoveries are a 

projections on total copper. Acid 

cure was for 5 days before leaching. 

Test period was 48 days Acid 

consumption "30lbs/ton ore". Cure 

time was 7 days 

Test period was 48 days Acid 

consumption "30lbs/ton ore". Cure 

time was 7 days 

Head grade is average of calculated 

head grades. Acid consumption =lbs 

acid /ton divided by lbs recovered Cu 













Mixed Copper Ores. Head Grade is 

calculated and averaged. 

Predominantly chrysocolla 

mineralization. Head grade is 

calculated and averaged 

Recovery = % of ASCu; acid 

consumption excludes EW Credit 

Recovery = % of ASCu; acid 

consumption excludes EW Credit 

Rock mineralogy black Cu WAD in 

quartz monzonite, andesite, limonite 

and quartz monzonite, quartz 

monzonite porphyry and limonite or 

quartz monzonite and andesite. 

Samples were chrisocolla with minor 

amounts of malachite in quartz 

monzonite, quartz monzonite and 

limonite, or andesite. 

Sampes were limonite in quartz 

monzonite



14 MINERAL RESOURCE ESTIMATES 

14.1 INTRODUCTION 

An updated resource model has been prepared for the MacArthur Copper Project, located near 

Yerington Nevada, that supersedes previous estimates reported in the January 2011 Technical 

Report (Tetra Tech, 2011). The previous report included data obtained only through year 2010. 

Updated mineral resource estimates have been generated by incorporating new exploration 

drilling and sampling conducted as part of the 2011 exploration program conducted by Quaterra 

Alaska. A listing of all the drill holes used in the model is found in Appendix E, which includes 

the drill hole type and location. Interpolation characteristics have been defined based on the 

geology, drill hole spacing and geostatistical analysis of the data. The mineral resources have 

been classified by their proximity to the sample locations and are reported, as required by NI 43- 

101 guidelines, according to the CIM standards on Mineral Resources and Reserves. 

A total of 151 drill holes totaling 80,800 feet were added to the database used for the resource 

estimation. These included two holes for which data was unavailable at the time of the last 

estimate, but did not include three 2011 holes which were outside the model limits. This model 

differs from the previous resource estimate in the following ways: 

• The physical dimensions of the overall block model were increased to include the new 

drilling. The 2010 model (Tetra Tech, 2011) contained only 512 blocks in the x-direction 

whereas the current updated model contains 548. The other dimension of 400 blocks in 

the y-direction and 150 levels are unchanged. The individual block size utilized is 

25x25x20 feet. 

• The interpretation of the mineralized zones for the oxide, mixed (transition) and sulfide 

mineralization was updated based on the assay data obtained in 2011. 

• Dikes were added to the model, steeply dipping to the north, and modified search 

conditions were used in grade estimations. 

• Indicator kriging (IK), based on total copper grades above and below 0.12%, was 

employed to modify search conditions used in grade estimation. 

• More codes were assigned to blocks based on conditions of oxide-mixed-sulfide 

mineralization, the IK grade envelopes, and whether a block was within a dike or not. 

• Dynamic kriging was used to alter the direction of the search ellipsoid for each block 

sub-areas such as the SE-pit area and areas north and south of the “Hinge Line” that had 

distinct azimuth and directions of the search ellipsoids 

• Search parameters such as the maximum number of samples per sector, the number of 

samples per drill hole and the minimum samples required were modified. 

• Compositing was changed from 20 foot-level style to a 10 foot-zone style. 

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• Correlograms were used to model the spatial structure used in kriging. 

• Some modifications to the search parameters were done on criteria which determined 

whether an estimated block was to be classified as measured, indicated or inferred. 

• New jackknife studies were done to determine the required kriging parameters for block 

class. 

14.2 MACARTHUR RESOURCE ESTIMATION 

This section describes the methodology used in developing the mineral resource estimate for 

contained copper resources in the MacArthur deposit. Recent drilling on the MacArthur property, 

which further defines a significant amount of copper, coupled with updated geologic and mineral 

zone interpretations, provides the basis for an updated mineral resource estimate. Figure 14-1 

details the drill holes used in the updated estimation of the MacArthur deposit. 

The MacArthur mineral resource estimate was prepared in the following manner: 

• Data from an additional 151 holes was added for this report. 

• The density values for each rock code based on the previous studies are unchanged from 

the previous model. 

• The resource estimate was broken into two areas: the southeast historical pit area 

(variously called SE or SE-Pit area in this report) and the northwest area (variously called 

NW or NW-Out in this report). 

• Quaterra provided cross-sections with interpreted geology, lithology units, mineral zones 

(MinZones) and dikes. The MinZones were digitized by Quaterra and Tetra Tech (Tt) to 

produce wireframes surfaces. 

• Dike intercepts were used to create dike blocks, oriented east-west or N70 o W, dipping 

60 o to the north to allow separate grade interpolation within those blocks. 

• Statistics for drill hole five-foot interval assays were analyzed for each of the MinZone 

codes broken out by the southeast and northwest areas and by drill holes completed by 

Metech and Quaterra. 

• The interval assays were composited to a ten-foot zone-length. Statistics for the 

composites were analyzed for each of the rock codes within the southeast and northwest 

areas. As with the five-foot interval data, analyses were done separately on the Metech 

(Anaconda) and Quaterra data. 

• Geostatisitcal analysis was done on the ten-foot composite data. Unitized General 

Relative variograms (UGR) were generated. The directional variograms were modeled 

with the spherical function using a nugget and up to three nested structures. 

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• The quality of the variogram models was checked using a model-validation technique 

called “jackknifing”. The method helps determine the best variogram parameters to be 

used for the theoretical model, and to determine the best kriging parameters (range, 

direction and search parameters). 

• The resource model used multiple pass ordinary kriging (OK) to estimate total copper 

within each MinZone. The kriged grades were checked by comparing block, composite 

and assay histograms. 

• The block model values were visually inspected in multiple sections and plan maps. 

These values were compared to the drill hole traces that contain both interval assay data 

and composite data; 

• A resource classification of measured, indicated and inferred was developed based on a 

combination of minimum required data points, jackknifing and kriging error analysis. 

• The MacArthur copper resource was tabulated for volume, tonnage and contained metal 

for the measured, indicated and inferred classes. 

• The resource estimate was broken into two areas for evaluation: the southeast historical 

pit area and the northwest area. 

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Figure 14-1: Drill Location and Search Zones for the MacArthur 2011 Model 

14.3 MACARTHUR BLOCK MODEL 

Block model parameters for the MacArthur Copper Project were defined to best reflect both the 

drill spacing and current geologic interpretations. Table 14-1 shows the MacArthur block model 

parameters. 

Table 14-1: MacArthur Model Parameters 

MacArthur East Model Parameters X (Columns) Y (Rows) Z (Levels) 

Origin (lower left corner): 2,429,300 14,685,800 2,800 

Block size (feet) 25 25 20 

Number of Blocks 548 400 150 

Rotation 0 degrees azimuth from North to left boundary 

Composite Length 10 feet (Bench) 

An Excel database provided by Quaterra contained the pertinent drill hole and assay information 

for the MacArthur Copper deposit. The database contained 737 drill holes of which 676 drill 

holes from Quaterra and Anaconda (Metech) were used. The 61 holes removed included holes 

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with limited or no information on the assays (Pangea Gold 1991, Superior, USBM 1952, 

Anaconda 1955-57), and six Quaterra holes which were outside the model limits. Of the 676 

holes used, there are 280 Anaconda (Metech) RC holes and 396 Quaterra holes (58 core and 338 

RC holes). These drill holes traversed 257,895 feet, producing 51,258 total copper sample assay 

values at a nominal five feet in length. A list of drill holes used in this resource estimate is 

provided in Appendix E. 

Table 14-2 shows the MinZone codes, which can be considered levels of oxidation from 

topography changing with depth. Ideally, the top zone is the oxide zone with the chalcocite mix 

at a deeper level until a sulfide zone is encountered at depth. 

These zones were modeled as strata determined by Quaterra geologists by inspecting the 

mineralogy of samples from core and RC cuttings. The transition from air (MinZone 0) to the 

oxide zone/chalcocite mix transition was modeled as MinZone 10. The transition from the oxide 

zone to the sulfide was modeled as MinZone 20. The MinZone code below the chalcocite to 

sulfide zones was given the code MinZone 30. Finally, any undefined zones were given the code 

9999. 

By creating and then combining boundary lines on sections, these transition lines were used to 

generate MinZone transition surfaces. Then by using wireframe techniques the model produced 

3-D MinZone volumes (Tetra Tech used MicroModel ® and Quaterra used DataMine ® ). These 

initial zones codes were modified by the addition of 100 if the material was within a dike and the 

addition of 1 if indicator kriging defined the material to be within a higher grade zone. 

Table 14-2: MinZone Codes and Density 

MinZone Code Description Density (cu.ft/ton) 

0 Air and previously mined pit Air (0) and Mined (12.5) 

5, 6, 105, 106 Alluvium 12.5 

10, 11, 110, 111 Oxide zone 12.5 

20, 21, 120, 121 Chalcocite mix zone 12.5 

30, 31, 130, 131 Sulfide zone 12.5 

9999 Undefined 12.5 

Table 14-3 shows the count of the described MinZones of the 5-foot intervals. The table is 

broken into two parts. Note that the term “POLYGON” which designates a subset of the drill 

holes that has been isolated for statistical work. In the second section of Table 14-3 the drill hole 

data from the “NW-Out Area” (Northwest area) has been segregated for the count. Some of the 

counted assays are now above the current post-mine topography and are coded with a MinZone 

Code of 0. Even though these particular samples are above the current topography their assay 

values contain geostatistical information that was used in estimating remaining resources. 

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- 

Table 14-3: MinZone Interval Data Count and Drill hole Assay Statistics 

Quaterra-RC-2009, Quaterra-Core-2009, Metech-RC, Metech-RC-Twin, QM-Core-Twin-2010 

QM-RC-Twin-2010, Quaterra-RC-2010, Quaterra-Core-2010, QM_2011 

--------------------------------------------------------------------------------------- 

POLYGON: None 

ALL Areas 

TOTAL DRILL HOLES 676 

TOTAL LENGTH 257895.1 

EASTING NORTHING ELEVATION AZIMUTH DIP DEPTH 

MINIMUM 2430272.0 14686098.0 4563.0 0.0 45.0 0.0 

MAXIMUM 2442063.2 14694670.0 5491.0 357.1 90.0 2685.5 

AVERAGE 2437574.3 14689347.1 4818.6 64.1 75.3 381.5 

RANGE 11791.2 8572.0 928.0 357.1 45.0 2685.5 

AVERAGE VALUES OF SELECTED DATA 

LABEL NUMBER AVERAGE STD DEVIATION MIN. VALUE MAX. VALUE # MISS. 

FROM-TO 51550 5.00202 0.96702 0.40000 55.90002 0 

Cu% 51258 0.12289 0.19754 0.00000 13.80000 292 

asCu% 36206 0.03203 0.08982 0.00100 4.30000 15344 

QltCu% 18388 0.05205 0.13421 0.00500 6.16000 33162 

cnCu% 578 0.05150 0.14173 0.00500 2.43000 50972 

---------------------------------------------------------------------------------------- 

POLYGON: SE-PIT 

INSIDE SE-PIT AREA DRILL HOLES 405 




FROM-TO 21006 5.00028 0.75582 0.40000 54.80000 0 

Cu% 20885 0.17185 0.17696 0.00500 3.84000 121 

asCu% 9615 0.05856 0.12198 0.00300 2.30000 11391 

QltCu% 4202 0.07795 0.12791 0.00500 2.30000 16804 

cnCu% 180 0.02864 0.05539 0.00500 0.42000 20826 

--------------------------------------------------------------------------------------- 

POLYGON: NW-OUT 

OUTSIDE PIT AREA DRILL HOLES 271 




FROM-TO 30544 5.00323 1.08874 0.40001 55.90002 0 

Cu% 30373 0.08927 0.20380 0.00000 13.80000 171 

asCu% 26591 0.02244 0.07252 0.00100 4.30000 3953 

QltCu% 14186 0.04437 0.13508 0.00500 6.16000 16358 

cnCu% 398 0.06183 0.16573 0.00500 2.43000 30146 

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MinZone Interval Data Count (continued) 

Quaterra-RC-2009, Quaterra-Core-2009, Metech-RC, Metech-RC-Twin, QM-Core-Twin-2010 

QM-RC-Twin-2010, Quaterra-RC-2010, Quaterra-Core-2010, 

POLYGON: None 

Drill Data before 2011 


MINIMUM 2430272.0 14684114.0 4533.3 0.0 45.0 35.0 

MAXIMUM 2442821.8 14694670.0 5491.0 357.1 90.0 2000.0 

AVERAGE 2437692.1 14689147.7 4801.0 55.3 78.3 338.3 

RANGE 12549.8 10556.0 957.7 357.1 45.0 1965.0 

TOTAL COUNT 531 




FROM-TO 17682 5.01859 1.39921 0.40001 55.90002 0 

Cu% 17547 0.09064 0.19562 0.00000 8.85000 135 

asCu% 16326 0.02135 0.06956 0.00100 3.56000 1356 

QltCu% 3979 0.06491 0.15691 0.00500 4.81000 13703 

cnCu% 398 0.06183 0.16573 0.00500 2.43000 17284 

--------------------------------------------------------------------------------------- 

QM_2011 

POLYGON: None 

Drill Data 2011 


MINIMUM 2431342.5 14686098.0 4613.9 0.0 45.0 0.0 

MAXIMUM 2440325.0 14694458.0 5301.7 270.0 90.0 2685.5 

AVERAGE 2437248.7 14689948.0 4876.0 92.3 64.7 535.1 

RANGE 8982.5 8360.0 687.7 270.0 45.0 2685.5 

TOTAL COUNT 151 




FROM-TO 12862 4.98220 0.34916 0.59998 20.00000 0 

Cu% 12826 0.08739 0.21453 0.00000 13.80000 36 

asCu% 10265 0.02417 0.07696 0.00100 4.30000 2597 

QltCu% 10207 0.03636 0.12464 0.00500 6.16000 2655 

14.4 ASSAY DATA 

The assay data was assigned MinZones using the interpreted wireframes. The process first used 

DataMine ® to assign a MinZone to each 25x25x20-foot block within the model specified in 

Table 14-1. When the majority of a block fell within the interpreted MinZone wireframe it was 

assigned the corresponding code. These coded blocks were then imported into MicroModel ® and 

used to “back-mark” each sample using a simple majority rule. Table 14-4 gives the count of 

MinZone for composites. The table is divided into three sections. The first section gives a count 

of the Minzone codes for assays from all drill hole types, and no limiting polygon. For example 

the count of assays above the current topography (Code 0) is 7,733. The second section gives the 

statistics for the assays. For example the mean Cu grade in % in MinZone 0 is 0.327. The 

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average for all zones, combined SE and SW areas is 0.125. The third section is a classic 

histogram plotted with a log scale. 

MINZONE COUNT FOR SAMPLES 

Table 14-4: Statistics of Cu Assay Data (All Areas) 

All Classes: 1 = Quaterra RC 2 = Quaterra Core 3 = Metech RC 4 = Metech RC-Twin 

5 = QMT-Core-Twin 6 = QMT-RC-Twin 11 = quaterra-2010 13 = quaterra-core-2010 14 = QM- 

2011 

POLYGON LIMITING FILE USED: None 

CODE* COUNT MINCOL MAXCOL MINROW MAXROW MINLEV MAXLEV 

0 7733 39 507 12 355 1 130 

5 701 44 511 32 253 81 135 

6 95 167 426 55 231 89 117 

10 14118 42 511 12 269 68 134 

11 10153 50 491 30 250 71 130 

20 5343 39 511 12 307 70 118 

21 2622 82 462 51 253 70 119 

30 3624 102 455 52 253 47 114 

31 487 158 441 72 253 48 108 

105 39 122 439 90 231 89 124 

106 16 273 302 210 232 102 108 

110 2468 42 474 52 269 75 126 

111 1167 119 431 35 233 74 112 

120 1257 39 474 39 251 73 114 

121 537 119 437 72 250 74 115 

130 939 122 455 72 251 58 114 

131 43 261 400 191 253 68 83 

9999 208 184 425 307 347 1 1 

TOTAL 51550 

• 2010 43-101 base codes of 5, 10, 20 and 30 have been modified such that: 

All Dike Material has had a 100 added to the base code. 

All composites within a grade shell of 0.12% Cu have a 1 added to the base code. 

Codes 0 and 9999 are exceptions. 

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All Cu Assay Statistics (continued) 

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Table 14-5 and Table 14-6 show the statistics in the SE area and NW areas respectively. 

The SE data is “lognormal-like”, in that it generally follows a bell shaped curve with some 

notable deviations and an average total copper grade of 0.172. The NW data has a mean grade of 

0.091 with a highly skewed distribution. 

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Table 14-5: SE-Pit Area Cu Assay Statistics 

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Table 14-6: NW Area TCu Assay Statistics 

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Figure 14-2 shows the SE and NW copper assay grades together in a side-by-side format. The 

length of the histogram bars is proportional to the total count of assays from each area. Note the 

enhanced grade of the SE-Pit area (blue bars) is quite apparent with respect to the NW-Out 

values (green bars). 

Figure 14-2: Side-by-Side Histograms – TCu% Assay SE-PIT area and NW-OUT area 

14.5 COMPOSITE DATA 

The assay data was composited using a 10-foot “zone method”. The zone method is a variant of 

down hole compositing, with the distinction that the composite begins as the drill interval enters 

a rock code zone. This method tends to reduce averaging composites across zones. The process 

first used DataMine ® to assign a MinZone to each 25x25x20-foot block within the model 

specified in Table 14-1. When the majority of a block fell within the interpreted MinZone 

wireframe it was assigned the code. These coded blocks were then imported into MicroModel ® 

and used to “back-mark” each composite using a simple majority rule. No capping was applied. 

Table 14-7 gives the count of MinZone for composites. Note that the plotted histograms shown 

in Table 14-8 through Table 14-10 are more lognormal-like than the original assay data. Also 

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note that the average values of the composites are quite similar to the averages shown for assays, 

and the coefficient of variation (CV) has been reduced. 

Table 14-7: MinZone Composite Count (All Areas) 

MINZONE COUNT FOR COMPOSITES (10-foot Zone) 

Quaterra & Metech Class: 1 = Quaterra RC 2 = Quaterra Core 3 = Metech RC 4 = Metech RC-Twin 

5 = QMT-Core-Twin 6 = QMT-RC-Twin 11 = quaterra-2010 13 = quaterra-core-2010 14 = QM-2011 

POLYGON LIMITING FILE USED: None 

CODE* COUNT MINCOL MAXCOL MINROW MAXROW MINLEV MAXLEV 

0 560 223 421 87 250 91 108 

5 347 44 541 32 253 81 135 

6 47 167 426 55 231 90 117 

10 7118 42 541 12 269 68 134 

11 5084 50 541 30 250 72 130 

20 2669 39 541 12 307 70 118 

21 1317 82 462 51 253 70 119 

30 1816 102 455 52 253 47 114 

31 244 158 441 72 253 48 108 

105 15 122 437 90 209 90 124 

106 9 273 302 210 232 102 108 

110 1227 42 474 52 269 75 126 

111 565 119 431 35 233 74 112 

120 624 39 474 39 251 73 114 

121 268 119 437 72 250 74 115 

130 525 122 455 72 251 66 114 

131 18 261 400 191 253 68 83 

9999 3577 39 541 1 355 1 130 

TOTAL 26030 

• 2010 43-101 base codes of 5, 10, 20 and 30 have been modified such that: 

All Dike Material has had a 100 added to the base code. 

All composites within a grade shell of 0.12% Cu have a 1 added to the base code. 

Codes 0 and 9999 are exceptions. 

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Table 14-8: All Cu Assay Statistics for Quaterra Composites 

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Table 14-9: SE Area Cu Assay Statistics for Quaterra Composites 

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Table 14-10: NW Area Cu Assay Statistics for Quaterra Composites 

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Figure 14-3 shows the SE and NW copper composite grades together in a side-by-side format. 

The length of the histogram bars is proportional to the total count of assays from each area. 

Again the enhanced grade of the SE-Pit area (blue bars) is quite apparent with respect to the 

NW-Out values (green bars). 

Figure 14-3: Side-by-Side Histograms – TCu% Composites SE-PIT area & NW area 

14.6 GEOSTATISTICAL ANALYSIS AND VARIOGRAPHY 

A total of twenty-two (21 directional and one omni-directional) variograms were calculated 

using MicroModel® for each MinZone within each area. The program searches along each 

direction for data pairs within a 12.5-degrees window angle and 5-feet tolerance band. All 

experimental variograms are inspected so that spatial continuity along a primary, secondary and 

tertiary direction can be modeled. 

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Each variogram model was then validated using the “jackknifing” method. This method 

sequentially removes values and then uses the remaining composites to krige the missing value 

using the proposed variogram. 

Figure 14-4 shows the horizontal omni-variogram of indicator 0-1 values of total copper grades 

equal to and below 0.12% Cu (indicator=0) and above (indicator=1). A nugget and three 

spherical model structures are modeled. Indicator kriging was used to partition the blocks into 

“high (1)” and “low (0)” grade sub-areas. These sub-areas were used to recode the blocks. For 

example, a block with a 10 code within a high grade area was recoded with an “11” code. 

The second panel of Figure 14-5 shows two figures containing experimental correlograms of 

total copper in various directions. A correlogram can be considered as a variation of a variogram, 

with the graph plotting the correlation of data at increasing separation distances. A perfect 

correlation plots as a “1.0”. A nugget effect is when two samples at nearly the same location 

have a correlation less than 1. When samples are no longer correlated (0.0), the distance is 

considered to be the “range” of the data. 

Figure 14-4: 0.12% Indicator Variograms (Omni Direction) For NW-Out and SE-Pit Areas 

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(a) SE Correlograms 

(b) NW-OUT Correlogram 

Figure 14-5: Selected Cu% Correlograms For SE-Pit And NW-Out Areas 

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14.7 KRIGING 

Kriging requires not only a variogram model but other search parameters. Table 14-11 shows the 

search parameters and variogram parameters used for block kriging of total copper. As 

discussed, the initial MinZone codes of 10, 20 and 30 have been modified. Zone codes within 

modeled dikes have been increased by 100. Zones within areas considered to be of higher grade 

using a 0.12% indicator have been increased by 1. Dynamic kriging was used in this estimate; 

the method changes both the search and variogram parameters for every estimate block. The dip 

parameter for all 30 and 100 codes are defined in the table. 

Table 14-11: Variogram and Search Parameters 

For example, for MinZone 10 in the SE area, the search ellipse of 300x400x100 feet will be 

oriented so its primary axis has an azimuth of 20 degrees north, a dip of 0 degrees. Conditions 

regarding the minimum number of drill holes to be used for each resource class and how an 

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additional condition that the kriging error must be within certain bounds may also impact a 

resource classification are discussed further in Section 14.8. 

Table 14-12 gives the count of potentially estimated blocks for each of the MinZones in the SE 

and NW areas. It should be noted that not all of these blocks will be estimated. Table 14-13 

through Table 14-15 give the statistics for the kriged blocks within the SE and NW Areas, 

respectively. 

Figure 14-6(a) shows the block values in the SE-Pit area (Zones 10 and 20) for measured + 

indicated (M&I, green bars) and inferred (red bars) in side-by-side format. Note that M&I have 

sharply defined the higher grade population. The inferred distribution is more complex, perhaps 

reflecting a mix of several grade populations. Figure 14-6(b) shows a lesser grade enhancement 

in the NW zones 10 and 20 for the M&I blocks versus the inferred. In both cases, the 

enhancement is due to the geologic modeling. 

Table 14-12: MinZone Block Count (All Areas) 

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Table 14-13: SE-Pit and NW Areas Cu Block Statistics 

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Table 14-14: SE Area Cu Block Statistics 

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Table 14-15: NE Area Cu Block Statistics 

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(a) INF (RED) and M&I (Green) Blocks – SE-Pit Area 

(b) INF (RED) and M&I (Green) Blocks – NW-Out Area 

Figure 14-6: Side-by-Side Histograms M&I vs INF for (a) SE and (b) NW-Out 

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14.8 KRIGING ERROR AND RESOURCE CLASSIFICATION 

Tt used a two-part approach to classify the total copper resources. This approach takes into 

account the spatial distribution of the drilling, the distance to the nearest data points used to 

estimate a block, and finally the relative kriging error generated by the estimate. Tt has found 

this approach to be very robust and provide highly reproducible results. The following points 

detail this approach. 

• A measured block requires 16 samples, with a maximum of five samples per sector in a 6 

sector search pattern and a maximum of 2 composites coming from a single drill hole. 

This implies that in most cases, for a block to be classified as measured there must be a 

least 8 drill holes in four cardinal directions 

• The constraints for an indicated block are not as stringent as those for a measured block. 

An indicated block requires a minimum of 6 samples, with a maximum of 4 samples per 

sector in a sector search pattern and a maximum number of 2 samples coming from a 

single drill hole. This implies that for most cases an indicated block must have at least 3 

drill holes in three of the four cardinal directions. 

• Relaxing the constraints even more, an inferred block requires a minimum of 2 samples, 

with a minimum of 2 samples per sector and a maximum of 2 composites from a single 

drill hole. This implies that an inferred block must have a least one drill hole from one of 

the four cardinal directions. 

In addition to the search parameters, kriging error comes into play when determining if a block 

falls into a particular class. Tt has found that by plotting the kriging error as a log-probability 

plot, there is a natural break in the distribution which signifies when the error is too great to 

allow a block to be classified as measured or indicated (Figure 14-7). In the case of the 

MacArthur deposit, any block with a kriging error of 0.75 or larger was classified as inferred. 

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Figure 14-7: Probability plot of kriging error 

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(a) Scatter plot of the Jackknife estimate and target grades 

(b) Histogram of the Jackknife estimate and target difference. 

Figure 14-8: Jackknife Method of Model Validation 

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The use of a model validation technique called “jackknifing” has been used to help validate the 

chosen search parameters. The technique removes, in sequence, a target value and uses the 

chosen estimation method to predict its value. The target and estimate are then compared. Figure 

14-8(a) shows a scatter plot of the plotted target and estimated grades. A perfect estimation 

would produce a 45-degree slope of points and a correlation of 1.0. This plot has a correlation of 

0.62 with approximately 80% of the points falling within the plotted ellipse. Figure 14-8(b) show 

the histogram of the difference between the target and estimate grades. An unbiased estimation 

would show a difference of zero. A precise estimate should have a small spread in the 

differences. Figure 14-9 show a jackknife scatterplot with the results of the three passes for 

MinZones 10 and 11. Correlation for pass 1 is 0.7. Pass 2 has a correlation of 0.5 and Pass 3 a 

correlation of 0.3. The nested ellipses capture approximately 80% of the plotted points for each 

test. 

Figure 14-9: Jackknife validation of kriging model (SE Area, MinZones 10 and 11) 

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14.9 VALIDATION OF BLOCK MODEL: VISUAL AND STATISTICAL CHECKS 

The resultant block model was validated both with visual checking of the block model grades 

against drill hole assays and composites using DataMine ® . Figure 14-10 shows a graphic of one 

of the statistical comparisons. Side-by-side histograms of samples assays, composite grades and 

block grades shows that the three histograms are well centered. The histogram of the assays 

shows the grade spikes at low grades, which may be a function of laboratory detection limits. 

Figure 14-10: Side-by-Side Samples, Composites and Blocks 

Figure 14-11 and Figure 14-12 show an east-west cross section used for visual check of the 

copper grades and resource classes and Figure 14-13 and Figure 14-14 similarly show a northsouth 

section used for visual inspection. Surfaces shown below the current topography are the 

bottom of oxide /top of chalcocite, and the bottom of chalcocite-mix / top of primary sulfides. 

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Figure 14-11: East West Cross Section Looking North (Cu blocks) 

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Figure 14-12: East-West Cross Section Looking North (Resource Class) 

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Figure 14-13: North-South Cross Section Looking West (Cu Blocks) 

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Figure 14-14: North South Cross Section Looking West (Resource Class) 

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14.10 MINERAL RESOURCE ESTIMATE 

A summary of the measured copper resource is shown in Table 14-16. A summary of the 

indicated copper resources is shown in Table 14-17. The combined Measured and Indicated 

Copper Resources are shown in Table 14-18, and a summary of the inferred copper resources by 

deposit area is shown in Table 14-19. The base case cutoff grade for the leachable resources is 

0.12% Cu. The base case cutoff grade for the primary sulfide resources is 0.15% Cu. Both of 

these values are representative of actual operating cutoff grades in use as of the date of this 

report. It is Tt’s opinion that the MacArthur Mineral Resources meet the current CIM definitions 

for classified resources. 

A “Measured Mineral Resource” is that part of a Mineral Resource for which quantity, 

grade or quality, densities, shape, and physical characteristics are so well established that 

they can be estimated with confidence sufficient to allow the appropriate application of 

technical and economic parameters, to support production planning and evaluation of the 

economic viability of the deposit. The estimate is based on detailed and reliable exploration, 

sampling and testing information gathered through appropriate techniques from locations 

such as outcrops, trenches, pits, workings and drill holes that are spaced closely enough to 

confirm both geological and grade continuity. 

Mineralization or other natural material of economic interest may be classified as a Measured 

Mineral Resource by the Qualified Person when the nature, quality, quantity and distribution of 

data are such that the tonnage and grade of the mineralization can be estimated to within close 

limits and that variation from the estimate would not significantly affect potential economic 

viability. This category requires a high level of confidence in, and understanding of, the geology 

and controls of the mineral deposit. 

An “Indicated Mineral Resource” is that part of a Mineral Resource for which quantity, 

grade or quality, densities, shape and physical characteristics, can be estimated with a level 

of confidence sufficient to allow the appropriate application of technical and economic 

parameters, to support mine planning and evaluation of the economic viability of the 

deposit. The estimate is based on detailed and reliable exploration and testing information 

gathered through appropriate techniques from locations such as outcrops, trenches, pits, 

workings and drill holes that are spaced closely enough for geological and grade continuity 

to be reasonably assumed. 

Mineralization may be classified as an Indicated Mineral Resource by the Qualified Person 

when the nature, quality, quantity and distribution of data are such as to allow confident 

interpretation of the geological framework and to reasonably assume the continuity of 

mineralization. The Qualified Person must recognize the importance of the Indicated Mineral 

Resource category to the advancement of the feasibility of the project. An Indicated Mineral 

Resource estimate is of sufficient quality to support a Preliminary Feasibility Study which can 

serve as the basis for major development decisions. 

An “Inferred Mineral Resource” is that part of a Mineral Resource for which quantity and 

grade or quality can be estimated on the basis of geological evidence and limited sampling 

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and reasonably assumed, but not verified, geological and grade continuity. The estimate is 

based on limited information and sampling gathered through appropriate techniques from 

locations such as outcrops, trenches, pits, workings and drill holes. 

Due to the uncertainty that may be attached to Inferred Mineral Resources, it cannot be assumed 

that all or any part of an Inferred Mineral Resource will be upgraded to an Indicated or 

Measured Mineral Resource as a result of continued exploration. Confidence in the estimate is 

insufficient to allow the meaningful application of technical and economic parameters or to 

enable an evaluation of economic viability worthy of public disclosure. Inferred Mineral 

Resources must be excluded from estimates forming the basis of feasibility or other economic 

studies. 

Source: CIM DEFINITION STANDARDS - For Mineral Resources and Mineral Reserves CIM Standing 

Committee on Reserve Definitions, November 27, 2010 

Table 14-16: Measured Copper Resources 

MEASURED RESOURCES 


May 2011 

Oxide and Chalcocite Material 



(MinZone 30) 

Cutoff 

Grade 

Tons 

Average 

Grade 

Contained 


%TCu (x1000) %TCu (lbs x 1000) 

0.5 1,444 0.675 19,491 

0.4 3,196 0.548 35,041 

0.35 5,074 0.483 49,025 

0.3 8,633 0.417 71,930 

0.25 15,929 0.35 111,599 

0.2 33,472 0.283 189,518 

0.18 43,753 0.261 228,566 

0.15 58,388 0.237 276,993 

0.12 

0.5 

0.4 

0.35 

0.3 

0.25 

0.2 

0.18 

71,829 0.218 313,174 

0.15 

N/A N/A N/A 

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Table 14-17: Indicated Copper Resources 

INDICATED COPPER RESOURCES 


May 2011 




(MinZone 30) 

Cutoff 

Grade 

Tons 

Average 

Grade 

Contained 


%TCu (x1000) %TCu (lbs x 1000) 

0.5 1,957 0.753 29,484 

0.4 3,533 0.615 43,442 

0.35 5,018 0.543 54,516 

0.3 7,618 0.468 71,259 

0.25 13,930 0.379 105,478 

0.2 31,949 0.29 185,049 

0.18 45,554 0.26 236,607 

0.15 67,271 0.229 308,639 

0.12 87,264 0.208 362,320 

0.5 98 0.72 1,411 

0.4 193 0.586 2,263 

0.35 273 0.523 2,857 

0.3 354 0.478 3,382 

0.25 507 0.416 4,216 

0.2 670 0.369 4,938 

0.18 796 0.34 5,414 

0.15 1,098 0.292 6,408 

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Table 14-18: Measured + Indicated Copper Resources 

MEASURED + INDICATED RESOURCES 


May 2011 




(MinZone 30) 

Cutoff 

Grade 

Tons 

Average 

Grade 

Contained 


%TCu (x1000) %TCu (lbs x 1000) 

0.5 3,401 0.72 48,974 

0.4 6,730 0.583 78,485 

0.35 10,092 0.513 103,544 

0.3 16,251 0.441 143,171 

0.25 29,859 0.364 217,075 

0.2 65,421 0.286 374,601 

0.18 89,306 0.26 465,106 

0.15 125,659 0.233 585,822 

0.12 159,094 0.212 675,513 

0.5 98 0.72 1,411 

0.4 193 0.586 2,263 

0.35 273 0.523 2,857 

0.3 354 0.478 3,382 

0.25 507 0.416 4,216 

0.2 670 0.369 4,938 

0.18 796 0.34 5,414 

0.15 1,098 0.292 6,408 

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Table 14-19: Inferred Copper Resources 

INFERRED COPPER RESOURCES 


May 2011 




(MinZone 30) 

Cutoff 

Grade 

Tons 

Average 

Grade 

Contained 


%TCu (x1000) %TCu (lbs x 1000) 

0.5 4,294 0.657 56,423 

0.4 9,656 0.538 103,899 

0.35 15,357 0.477 146,444 

0.3 25,851 0.414 213,788 

0.25 43,695 0.356 311,108 

0.2 82,610 0.293 483,929 

0.18 109,920 0.267 587,412 

0.15 166,930 0.232 774,889 

0.12 243,417 0.201 979,510 

0.5 10,644 0.819 174,413 

0.4 18,442 0.653 240,742 

0.35 23,316 0.594 277,181 

0.3 33,831 0.511 345,415 

0.25 53,060 0.423 449,312 

0.2 89,350 0.341 609,188 

0.18 101,375 0.323 654,680 

0.15 134,900 0.283 764,074 

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15 MINERAL RESERVE ESTIMATES 

At this time, the MacArthur Copper Property does not have any CIM definable mineral reserves. 

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16 MINING METHODS 

Mining of the MacArthur deposit will be done by open pit methods utilizing a traditional drill, 

blast, load and haul sequence. Ore will be delivered to the run of mine heap leach and waste rock 

will be deposited in the waste dumps located to the north and south of the proposed pits along 

with pit backfilling of the pits later in the mine life. The pit design is based on a 20 foot bench 

height to match the resource model bench height. The mine plan calls for the delivery of 15 

million short tons (tons) per year (approximately 41,000 tons per day, tpd) of ore to the heap 

leach. During peak production about 96,000 tpd of total material (ore plus waste) will be mined. 

The mine equipment fleet requirements are estimated so as to mine and deliver the ore and waste 

tonnages to the appropriate locations. The major mining equipment fleet will include 9 inch blast 

hole drills, 26 cubic yard (cu yd) hydraulic shovels, a 17 cu yd loader, and multiple 150 ton 

trucks. From the estimate of the mine fleet requirements, an estimate of capital and operating 

costs was developed. 

The open pit resource tonnages included in this section are a sub-set of the mineral resource 

presented in Section 14. The open pit resource is contained within three pits. The mine schedule 

includes a brief pre-production period followed by 18 years of ore delivery to the heap leach pad 

totaling 271 million tons averaging 0.21% total copper. The life of mine average waste to ore 

ratio is 0.90. The heap leach material is the sum of oxide overburden, oxide rock and transition 

rock. Metallurgical test work estimates recovery of 70% for oxide material in the main pit, 65% 

for oxide in the north area and Gallagher and 60% recovery for all transition material from their 

respective total copper grades. No sulfide material is included in the open pit resource used for 

the mine production schedule. 

16.1 GEOTECHNICAL PARAMETERS 

No geotechnical investigations for pit slope angles have been completed for this PEA. An overall 

slope angle of 42 degrees was used for the pit definition floating cone runs and an inter-ramp 

slope angle of 45 degrees was used for the final pit and phase designs for the east, south and west 

pit walls. An inter-ramp slope angle of 46 degrees was used on the north wall because it is 

cutting the north dipping bedding. 

16.2 DILUTION MODELING AND FACTORS 

The resource model is described in section 14. At this time, no additional dilution factors or 

mining losses have been applied to the grade model. 

16.3 OPEN PIT MINING 

The PEA open pit design is based on a floating cone geometry using the available process 

recoveries, cost data and copper prices which range from $2.00 to $2.85/lb copper. Table 16-1 

summarizes inputs to the floating cone algorithm used for pit definition and internal phases. The 

process costs and recoveries are provided by M3 Engineering & Technology (M3). IMC 

provided the mining costs based on recent, similar size projects. Process cost for the heap 

leaching of copper is estimated on per pound of refined copper and the G&A costs are per ton of 

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ore. Mining costs are estimated using a base cost of $1.25 per ton of material moved with an 

addition haul cost from benches below the 4670 elevation of $0.02/t per 20 foot bench. The 

floating cones were run with a discount rate of 0.5% per bench of depth. 

Final pits for the PEA are designed from the floating cone geometry with the inclusion of haul 

ramps and the smoothing of the pit walls. The inter-ramp slope angle is 45 degrees on all but the 

north walls where it is 46 degrees. Ramps have a maximum grade of 10% and are 112 feet wide 

(including an allowance for berms and ditches). There are three final pits (Main, North and 

Gallagher).The Main pit is sub-divided into three mining phases. The North pit is divided into 

two phases with Gallagher a single phase. The mining phases and the copper price cone 

geometry used for each of the mining phases are in Table 16-2. The pits and phases are tabulated 

on Table 16-3 and illustrated as Figure 16-1 (final pits) and Figure 16-2 through Figure 16-7 

show the mining phase development. The pit exits are on the east side of the pits because the 

heap leach area is located to the east of the pits. On the figures, the bold lines show the mining 

faces of the current phase and the gray lines represent mining advances from previous mining 

phases. 

The tabulations on Table 16-3 show the heap leach tonnage split into material types and class 

(measured (26%), indicated (29%) and inferred (45%) categories, all of which will be used for 

developing the mine production schedule. The term ‘ore’ is used to describe the tonnage being 

delivered to the heap, but this does not imply that there is a mineral reserve at MacArthur. No 

pre-feasibility or feasibility study has been completed which is required to declare a mineral 

reserve. The tonnages of the heap leach material are tabulated using the 0.12% total copper 

cutoff grade. The mine schedule presented later in Table 16-4 uses an elevated cutoff grade in 

some years to improve the head grade, thus all of the tonnage shown on Table 16-3 will not be 

delivered to the heap for leaching. 

Table 16-1: Pit Definition Inputs 

Copper Recovery: 

Oxide in Main MacArthur Pit 70% of TCu 

Oxide in North Area and 65% of TCu 

Gallagher 

Transition ore type 60% of TCu 

Costs: 

Process (Heap and SXEW) $1.02/lb recovered copper 

G&A Cost $0.50/ton ore 

Ore Haul and Heap dozing $0.25/ton ore 

Mining (ore and waste), base $1.50/ton 

cost 

Mining Lift Charge $0.02/ton per 20 ft bench below 

4760 

Floating Cone Discount Rate 0.5% per 20 ft bench 

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Table 16-2: Floating Cone Geometries Used for Pit Designs 

Mining Location Active Mining Floating Cone Used to Design Phase 

Phase 

Years 

1 Main Pit PP – 5 $2.00/lb Cu cone 

2 Main Pit 3 – 10 Increment between $2.00/lb & $2.25/lb cone 

3 North Area 3 – 12 $2.50/lb cone 

4 North Area 8 – 15 Increment between $2.50/lb & $2.75/lb cone 

5 Gallagher 12 – 17 $2.60/lb cone 

6 Main Pit 12 - 19 Increment between $2.25/lb & $2.85/lb cone 

Table 16-3: Phase Tonnage and Grade Available for Mine Production Schedule 

Phase Class Overburden Oxide 

Transition Total Ore Waste Total Waste/Ore 

kton %Tcu kton %Tcu kton %Tcu kton %Tcu kton kton Ratio 

Measured 426 0.19 39,656 0.22 755 0.23 40,837 0.22 

Indicated 283 0.18 15,206 0.20 2,154 0.24 17,643 0.20 

1 Meas&Indic 709 0.19 54,862 0.21 2,909 0.24 58,480 0.22 

Inferred 26 0.19 186 0.18 2,020 0.22 2,232 0.22 

Total 735 0.19 55,048 0.21 4,929 0.23 60,712 0.22 9,905 70,617 0.16 

Measured 94 0.17 17,293 0.20 14 0.17 17,401 0.20 

Indicated 362 0.16 22,652 0.18 120 0.18 23,134 0.18 

2 Meas&Indic 456 0.16 39,945 0.19 134 0.18 40,535 0.19 

Inferred 333 0.17 6,387 0.19 2,537 0.19 9,257 0.19 

Total 789 0.17 46,332 0.19 2,671 0.19 49,792 0.19 9,574 59,366 0.19 

Measured 0 409 0.20 647 0.22 1,056 0.21 

Indicated 56 0.20 3,967 0.19 3,550 0.23 7,573 0.21 

3 Meas&Indic 56 0.20 4,376 0.19 4,197 0.23 8,629 0.21 

Inferred 604 0.19 21,229 0.19 16,678 0.24 38,511 0.21 

Total 660 0.20 25,605 0.19 20,875 0.24 47,140 0.21 55,932 103,072 1.19 

Measured 0 173 0.21 4,520 0.31 4,693 0.31 

Indicated 4 0.14 1,179 0.18 13,371 0.28 14,554 0.27 

4 Meas&Indic 4 0.14 1,352 0.18 17,891 0.29 19,247 0.28 

Inferred 112 0.15 4,414 0.17 9,958 0.25 14,484 0.22 

Total 116 0.15 5,766 0.17 27,849 0.27 33,731 0.26 62,063 95,794 1.84 

Measured 2 0.14 565 0.23 0 567 0.23 

Indicated 22 0.18 3,250 0.21 348 0.26 3,620 0.21 

5 Meas&Indic 24 0.18 3,815 0.21 348 0.26 4,187 0.22 

Inferred 340 0.17 18,851 0.19 12,444 0.21 31,635 0.20 

Total 364 0.17 22,666 0.20 12,792 0.22 35,822 0.20 25,104 60,926 0.70 

Measured 4 0.16 4,688 0.18 2,017 0.21 6,709 0.19 

Indicated 135 0.14 11,166 0.17 4,232 0.21 15,533 0.18 

6 

Meas&Indic 139 0.14 15,854 0.17 6,249 0.21 22,242 0.18 

Inferred 755 0.16 17,933 0.18 10,347 0.26 29,035 0.21 

Total 894 0.16 33,787 0.18 16,596 0.24 51,277 0.20 74,778 126,055 1.46 

Measured 526 0.19 62,784 0.21 7,953 0.27 71,263 0.22 

Indicated 862 0.17 57,420 0.19 23,775 0.26 82,057 0.21 

Total Pits Meas&Indic 1,388 0.17 120,204 0.20 31,728 0.26 153,320 0.21 

Inferred 2,170 0.17 69,000 0.19 53,984 0.24 125,154 0.21 

Total 3,558 0.17 189,204 0.19 85,712 0.24 278,474 0.21 237,356 515,830 0.85 

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GALLAGHER 

NORTH AREA 

MAIN MACARTHUR PIT 

Figure 16-1: Final Pits 

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Figure 16-2: Mining Phase 1 in MacArthur Pit 

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Figure 16-4: Mining Phase 3 in North Pit Area 

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Figure 16-5: Mining Phase 4 in North Pit Area 

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Figure 16-6: Mining Phase 5 (Gallagher Pit) 

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16.4 MINING SCHEDULE 

A mining schedule to deliver 15 million tpy to the heap was developed from the mining phases 

previously described. Mining starts in the Main pit (phases 1 and 2), progresses to the North pit 

(phases 3 and 4), then to Gallagher (phase 5) and returns to the outer limits of the Main pit 

during phase 6. The mine production schedule is presented in Table 16-4. The tonnage of leach 

feed mined by mining phase by year is summarized in Table 16-5. During years 3, 4, 6 and 7, the 

cutoff grade for tonnage going to the heap is raised above the 0.12% total copper cutoff in order 

to improve the head grade to the heap. Maps showing mine advance in the pits are included in 

Section 16.5. 

The percent of the heap leach tonnages in the measured plus indicated and the inferred categories 

are shown on Table 16-6. During years 1 through 4, over 90% of the heap leach tonnage is in the 

measured plus indicated categories. In years 5 and 6, this percentage drops to 77% and 72%. 

Table 16-4: Production Schedule 

Cutoff 

Pounds of Copper Ore Tonnage by Type 

Grade Ore Tonnage & Grade Main Pit Area Other Areas All Areas 

Year %Tcu Waste Total Contained Recoverable Oxide (ox + ovb) Oxide (ox + ovb) Mixed 

ktons %Tcu Avg recCu ktons ktons x 1000 x 1000 kt %Tcu kt %Tcu kt %Tcu 

Pre-Prod 0.12 399 0.24 0.17 101 500 1,920 1,344 399 0.24 0 0 

1 0.12 15,000 0.21 0.15 4,042 19,042 62,851 43,996 15,000 0.21 0 0 

2 0.12 15,000 0.24 0.17 2,174 17,174 71,590 50,113 15,000 0.24 0 0 

3 0.14 15,000 0.21 0.15 5,000 20,000 62,445 43,572 14,204 0.21 790 0.17 6 0.24 

4 0.14 15,000 0.21 0.15 5,000 20,000 62,785 43,745 14,228 0.21 538 0.19 234 0.22 

5 0.12 15,000 0.21 0.14 5,000 20,000 62,358 41,313 9,306 0.20 1040 0.19 4,654 0.23 

6 0.14 15,000 0.19 0.13 15,000 30,000 58,368 40,166 11,794 0.19 3149 0.21 57 0.19 

7 0.13 15,000 0.19 0.13 20,000 35,000 56,740 38,134 8,181 0.18 5723 0.20 1,096 0.20 

8 0.12 15,000 0.20 0.13 20,000 35,000 60,395 39,018 4,369 0.18 6360 0.19 4,271 0.24 

9 0.12 15,000 0.20 0.13 20,000 35,000 60,655 38,923 4,405 0.19 4771 0.18 5,824 0.23 

10 0.12 15,000 0.21 0.13 20,000 35,000 63,881 39,475 1,190 0.19 4337 0.16 9,473 0.24 

11 0.12 15,000 0.23 0.14 20,000 35,000 69,140 41,913 0 2383 0.18 12,617 0.24 

12 0.12 15,000 0.25 0.15 17,403 32,403 75,051 46,021 1,202 0.16 2841 0.21 10,957 0.27 

13 0.12 15,000 0.24 0.15 18,156 33,156 71,036 45,714 3,440 0.21 7862 0.21 3,698 0.32 

14 0.12 15,000 0.22 0.14 20,000 35,000 65,071 41,281 3,670 0.18 5124 0.18 6,206 0.27 

15 0.12 15,000 0.21 0.13 17,261 32,261 63,113 39,970 4,400 0.16 4093 0.17 6,507 0.27 

16 0.12 15,000 0.20 0.12 12,749 27,749 58,730 37,360 4,639 0.17 3108 0.18 7,253 0.22 

17 0.12 15,000 0.20 0.13 15,256 30,256 59,232 38,748 8,809 0.18 418 0.14 5,773 0.23 

18 0.12 15,000 0.18 0.12 7,612 22,612 54,907 35,830 8,489 0.17 0 6,511 0.2 

19 0.12 482 0.18 0.11 194 676 1,735 1,052 31 0.18 0 451 0.18 

Total 270,881 0.21 0.14 244,948 515,829 1,142,003 747,688 132,756 0.20 52,537 0.19 85,588 0.24 

0.90 

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Table 16-5: Ore Production Schedule by Mining Phase 

Summary Sum of Oxide and Transition Ore Types 

Cutoff 

Main Pit North Area Pit Gallager Pit 

Grade Total Ore Phase 1 Phase 2 Phase 6 Total - Main Pit 

Phase 3 Phase 4 Total - North Area Gallager, Phase 5 

Year %Tcu 

ktons %Tcu ktons %Tcu ktons %Tcu ktons %Tcu ktons %Tcu ktons %Tcu ktons %Tcu ktons %Tcu ktons %Tcu 

Pre-Prod 0.12 399 0.24 399 0.24 399 0.24 

1 0.12 15,000 0.21 15,000 0.21 15,000 0.21 

2 0.12 15,000 0.24 15,000 0.24 15,000 0.24 

3 0.14 15,000 0.21 12,764 0.21 1,446 0.22 14,210 0.21 790 0.17 790 0.17 

4 0.14 15,000 0.21 6,920 0.20 7,543 0.22 14,463 0.21 537 0.19 537 0.19 

5 0.12 15,000 0.21 8,727 0.22 5,233 0.19 13,960 0.21 1,040 0.18 1,040 0.18 

6 0.14 15,000 0.19 11,796 0.19 11,796 0.19 3,204 0.21 3,204 0.21 

7 0.13 15,000 0.19 8,182 0.18 8,182 0.18 6,818 0.20 6,818 0.20 

8 0.12 15,000 0.20 4,398 0.18 4,398 0.18 10,394 0.21 207 0.14 10,601 0.21 

9 0.12 15,000 0.20 4,963 0.19 4,963 0.19 9,418 0.21 618 0.16 10,036 0.21 

10 0.12 15,000 0.21 3,268 0.19 3,268 0.19 8,057 0.24 3,676 0.19 11,733 0.22 

11 0.12 15,000 0.23 0 4,006 0.23 10,994 0.23 15,000 0.23 

12 0.12 15,000 0.25 1,203 0.16 1,203 0.16 148 0.19 11,214 0.27 11,362 0.27 2,436 0.21 

13 0.12 15,000 0.24 3,533 0.21 3,533 0.21 3,614 0.32 3,614 0.32 7,853 0.21 

14 0.12 15,000 0.22 5,289 0.22 5,289 0.22 2,589 0.33 2,589 0.33 7,122 0.18 

15 0.12 15,000 0.21 6,435 0.21 6,435 0.21 818 0.25 818 0.25 7,747 0.21 

16 0.12 15,000 0.20 6,514 0.20 6,514 0.20 8,486 0.20 

17 0.12 15,000 0.20 12,821 0.19 12,821 0.19 2,179 0.20 

18 0.12 15,000 0.18 15,000 0.18 15,000 0.18 

19 0.12 482 0.18 482 0.18 482 0.18 

Total 270,881 0.211 58,810 0.218 46,829 0.193 51,277 0.195 156,916 0.203 44,412 0.214 33,730 0.255 78,142 0.232 35,823 0.201 

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Table 16-6: Ore Production Schedule by Mining Phase and Resource Classification 

Summary 

Sum of Oxide and Transition Ore Types (Measured + Indicated) 

Cutoff 

Main Pit North Area Pit Gallager Pit 

TOTAL PITS 

Grade Total Ore Phase 1 Phase 2 Phase 6 Total - Main Pit 

Phase 3 Phase 4 Total - North Area Gallager, Phase 5 

% of 

Year %Tcu Total 

ktons %Tcu ktons %Tcu ktons %Tcu ktons %Tcu ktons %Tcu ktons %Tcu ktons %Tcu ktons %Tcu ktons %Tcu ktons %Tcu 

Pre-Prod 0.12 399 0.24 396 0.24 396 0.24 396 0.24 99.2% 

1 0.12 15,000 0.21 14,891 0.21 14,891 0.21 14,891 0.21 99.3% 

2 0.12 15,000 0.24 15,000 0.23 15,000 0.23 15,000 0.23 100.0% 

3 0.14 15,000 0.21 12,761 0.21 1,186 0.23 13,947 0.21 1 0.15 1 0.15 13,948 0.21 93.0% 

4 0.14 15,000 0.21 6,747 0.20 7,129 0.22 13,876 0.21 8 0.29 8 0.29 13,884 0.21 92.6% 

5 0.12 15,000 0.21 6,815 0.22 4,665 0.19 11,480 0.21 42 0.23 42 0.23 11,522 0.21 76.8% 

6 0.14 15,000 0.19 10,029 0.19 10,029 0.19 791 0.22 791 0.22 10,820 0.19 72.1% 

7 0.13 15,000 0.19 7,425 0.18 7,425 0.18 1,419 0.19 1,419 0.19 8,844 0.18 59.0% 

8 0.12 15,000 0.20 4,061 0.18 4,061 0.18 1,868 0.20 22 0.14 1,890 0.20 5,951 0.19 39.7% 

9 0.12 15,000 0.20 2,895 0.19 2,895 0.19 1,647 0.20 100 0.21 1,747 0.20 4,642 0.19 30.9% 

10 0.12 15,000 0.21 384 0.17 384 0.17 1,677 0.22 931 0.20 2,608 0.22 2,992 0.21 19.9% 

11 0.12 15,000 0.23 0 715 0.23 6,168 0.24 6,883 0.24 6,883 0.24 45.9% 

12 0.12 15,000 0.25 38 0.15 38 0.15 7 0.19 7,823 0.29 7,830 0.29 273 0.21 8,141 0.28 54.3% 

13 0.12 15,000 0.24 923 0.20 923 0.20 2,529 0.33 2,529 0.33 1,426 0.23 4,878 0.27 32.5% 

14 0.12 15,000 0.22 1,794 0.18 1,794 0.18 1,474 0.36 1,474 0.36 927 0.18 4,195 0.24 28.0% 

15 0.12 15,000 0.21 2,332 0.17 2,332 0.17 197 0.27 197 0.27 771 0.22 3,300 0.19 22.0% 

16 0.12 15,000 0.20 2,713 0.17 2,713 0.17 714 0.22 3,427 0.18 22.8% 

17 0.12 15,000 0.20 7,307 0.19 7,307 0.19 76 0.21 7,383 0.19 49.2% 

18 0.12 15,000 0.18 7,117 0.18 7,117 0.18 7,117 0.18 47.4% 

19 0.12 482 0.18 17 0.20 17 0.20 17 0.20 3.5% 

Total 270,881 0.211 56,610 0.216 37,774 0.194 22,241 0.182 116,625 0.203 8,175 0.209 19,244 0.279 27,419 0.258 4,187 0.214 148,231 0.213 54.7% 

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Ore Production Schedule by Mining Phase and Resource Classification (Continued) 

Summary Sum of Oxide and Transition Ore Types (Inferred) 

Cutoff 

Main Pit North Area Pit Gallager Pit TOTAL PITS 

Grade Total Ore Phase 1 Phase 2 Phase 6 Total - Main Pit Phase 3 

Phase 4 Total - North Area Gallager, Phase 5 

% of 

Year %Tcu Total 

ktons %Tcu ktons %Tcu ktons %Tcu ktons %Tcu ktons %Tcu ktons %Tcu ktons %Tcu ktons %Tcu ktons %Tcu ktons %Tcu 

Pre-Prod 0.12 399 0.24 3 0.14 3 0.14 3 0.14 0.8% 

1 0.12 15,000 0.21 109 0.19 109 0.19 109 0.19 0.7% 

2 0.12 15,000 0.24 0 0 0 0.0% 

3 0.14 15,000 0.21 2 0.28 259 0.20 261 0.20 789 0.18 789 0.18 1,050 0.18 7.0% 

4 0.14 15,000 0.21 173 0.22 414 0.18 587 0.19 529 0.19 529 0.19 1,116 0.19 7.4% 

5 0.12 15,000 0.21 1,911 0.22 568 0.17 2,479 0.21 998 0.18 998 0.18 3,477 0.20 23.2% 

6 0.14 15,000 0.19 1,767 0.20 1,767 0.20 2,413 0.22 2,413 0.22 4,180 0.21 27.9% 

7 0.13 15,000 0.19 757 0.20 757 0.20 5,398 0.20 5,398 0.20 6,155 0.20 41.0% 

8 0.12 15,000 0.20 337 0.18 337 0.18 8,527 0.21 185 0.14 8,712 0.21 9,049 0.21 60.3% 

9 0.12 15,000 0.20 2,069 0.19 2,069 0.19 7,771 0.21 518 0.15 8,289 0.21 10,358 0.21 69.1% 

10 0.12 15,000 0.21 2,884 0.18 2,884 0.18 6,380 0.24 2,744 0.18 9,124 0.22 12,008 0.21 80.1% 

11 0.12 15,000 0.23 0 3,291 0.22 4,826 0.21 8,117 0.21 8,117 0.21 54.1% 

12 0.12 15,000 0.25 1,164 0.17 1,164 0.17 141 0.19 3,391 0.23 3,532 0.23 2,163 0.21 6,859 0.21 45.7% 

13 0.12 15,000 0.24 2,610 0.22 2,610 0.22 1,086 0.31 1,086 0.31 6,428 0.21 10,124 0.22 67.5% 

14 0.12 15,000 0.22 3,495 0.24 3,495 0.24 1,115 0.28 1,115 0.28 6,194 0.19 10,804 0.21 72.0% 

15 0.12 15,000 0.21 4,103 0.24 4,103 0.24 621 0.24 621 0.24 6,976 0.21 11,700 0.22 78.0% 

16 0.12 15,000 0.20 3,802 0.21 3,802 0.21 7,772 0.19 11,574 0.20 77.2% 

17 0.12 15,000 0.20 5,514 0.19 5,514 0.19 2,103 0.21 7,617 0.20 50.8% 

18 0.12 15,000 0.18 7,883 0.19 7,883 0.19 7,883 0.19 52.6% 

19 0.12 482 0.18 465 0.18 465 0.18 465 0.18 96.5% 

Total 270,881 0.211 2,198 0.217 9,055 0.190 29,036 0.207 40,289 0.203 36,237 0.215 14,486 0.222 50,723 0.217 31,636 0.198 122,648 0.208 45.3% 

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16.5 WASTE DUMPS 

Two exterior dumps and one backfill pit have been designed to hold the 245 million tons of 

waste rock. The exterior dumps are located to the north and south of the pits and the pit backfill 

is primarily in the north pit area with some extending to the west of the north pit. The pit backfill 

has the potential of sterilizing further transition and sulfide resources and this will be further 

evaluated during the pre-feasibility (PFS) stage of the project. No extensive condemnation 

drilling has been done in the north and south waste dump areas. Further optimization of the mine 

plan including waste rock management will be completed as part of the PFS. 

The dumps are designed using 20 foot contours with a setback between them so that the overall 

slope of the dump face is 2.5:1.0 (horizontal to vertical) to allow for either concurrent 

reclamation or reclamation of the dump faces at the end of mining. The overall slopes in areas 

with haul ramps for truck access to the upper lifts will be even flatter than 2.5:1.0. The dump 

locations relative to the final pit are shown on Figure 16-8. 

The average density of the waste tonnage is 12.5 cuft/t in place, dry. A 30% swell factor has 

been applied for determining the waste volume required to hold the waste tonnage. The average 

density in the dump volume is 16.25 loose cuft/t, dry. The tonnage placed in the waste dumps by 

year is shown on Table 16-7. Figure 16-9 through Figure 16-13 show the pit and dump advances 

through the first 10 years of mining. 

Table 16-7: Waste Tonnage by Source and Destination 

Year 

Source Phases - Waste ktons Waste Destinations 

1 2 3 4 5 6 Total South North North Pit Backfill 

Dump Dump SW area Expand 

PP 101 101 101 

1 4,042 4,042 4042 

2 2,174 2,174 2174 

3 2,866 350 1,784 5,000 3,216 1,784 

4 1,308 3,013 679 5,000 4,321 679 

5 1,316 2,236 1,448 5,000 3,552 1,448 

6 4,513 10,487 15,000 4,513 10,487 

7 993 19,007 20,000 993 19,007 

8 188 12,365 7,447 20,000 188 19,812 

9 387 7,127 12,486 20,000 387 19,613 

10 858 3,820 15,322 20,000 858 19,142 

11 1,780 17,968 251 19,999 12,402 7,597 

12 161 6,503 4,162 6,577 17,403 6,577 10,826 

13 911 6,143 10,891 17,945 5,446 7,054 5,446 

14 1,031 6,777 12,402 20,210 6,000 7,808 6,402 

15 394 4,002 12,865 17,261 4,396 12,865 

16 2,670 10,079 12,749 12,749 

17 1,099 14,158 15,257 15,257 

18 7,612 7,612 7,612 

19 194 194 194 

Total 11,807 12,538 58,658 62,062 25,104 74,778 244,947 42,368 85,232 56,823 60,525 

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Figure 16-8: Final Pit and Dumps (including pit backfill) 

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Figure 16-9: End of Year 1 

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16.6 MINING EQUIPMENT 

Mine equipment requirements were calculated based on the annual mine production schedule, 

the mine work schedule, and equipment shift production estimates. The size and type of mining 

equipment is consistent with the size of the project, i.e. peak run-of-mine material movements of 

35 million tons per year. 

A summary of the total mine fleet by year for the major mine equipment is shown in Table 16-8. 

There is sufficient equipment to perform the following duties: 

• Construct additional roads, after preproduction, as needed to support mining activity, 

including pioneering work necessary for mine and dump expansion. 

• Strip topsoil in advance of mining and dumping. 

• Mine and transport the ore to the heap leach pad. Mine and transport the waste material 

from the pit areas to the waste storage areas. 

• Maintain all the mine work areas, in-pit haul roads, waste storage areas, and external haul 

roads. 

• Build and maintain in pit and on dump drainage structures as required. 

Mine equipment requirements were not estimated for the following activities: 

• Construction of any major surface water diversion channels and settlement ponds and 

dams, other than the ditching and sedimentation ponds for the waste storage areas. 

• Construction of the shop area and plant area. 

• Preproduction road construction outside of the immediate mine area. 

• Contouring or reclamation of dumps at the end of the project. 

• Mine dewatering for slope stability. 

The mine equipment fleet calculations are based on two 12 hour shifts for 355 days per year (710 

operating shifts). The number of pieces of equipment is based on equipment productivity for 

projects of similar tonnage movements. Detailed equipment requirement calculations on a year 

by year basis have not been completed for the PEA. 

The truck haul routes and profiles were measured for ore to the heap and waste to the waste 

dumps or pit backfill areas. Truck cycles were simulated to determine the cycle times and tons 

hauled per truck shift. From this, the number of operating trucks required was determined. The 

reference to specific equipment vendors is intended only to associate these vendors with the size 

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of the equipment included for this PEA and is not intended to be a recommendation of a 

particular equipment vendor. 

The major mine equipment consists of 9 inch blast hole drills, 26 cubic yard hydraulic shovels, a 

17 cubic yard loader, and multiple 150 ton trucks plus major and minor support equipment. 

Years 7 through 11 have the peak tonnage movement at 35 million tons per year; during year 6 

the mine is ramping up to this capacity. One additional drill, shovel and 3 more trucks are added 

to the fleet in year 5 to handle the increased tonnage. 

16.7 MINE LABOR 

Table 16-8: Mine Equipment 

Equipment Initial Fleet Peak Fleet 

Mine Major Equipment: 

Yr -1 & 1 Start Yr. 6 

9 inch Blast Hole Drill 3 4 

26 cu yd Shovel 1 2 

17 cu yd Front End Loader 1 1 

150 t Haul Truck 8 11 

D10T Track Dozer 2 2 

16m Motor Grader 2 2 

777F Water Truck 

Mine Major Support Eqpt.: 

2 2 

988HH Wheel Loader 1 1 

385C Excavator 1 1 

D8T Dozer 1 1 

735 ATD Haul Truck 2 2 

CM 785 Rock Drill 1 1 

1 cum Backhoe Loader 

Support Equipment: 

Cable reeler, fuel & lube 

1 1 

trucks, cranes, flatbed trucks, 

tire handler, forklifts light 

plants, etc. 

1 lot 1 lot 

Mine communications & 

radios, dispatch system, 

survey equipment, safety 1 lot 1 lot 

equipment, engineering & 

geology supplies 

Mine personnel includes all salaried supervisory and staff people working in mine operations, 

maintenance, and engineering/geology departments, and the hourly people required to operate 

and maintain the drilling, blasting, loading, hauling, and mine support activities. In general 

mining activities end once the ore is delivered to the heap lead pad. 

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The mine operating and maintenance labor will operate on a three crew rotation. The estimates of 

personnel are based on similar size projects. The salaried staff includes supervision labor in 

operations and maintenance and the personnel in the engineering and geology departments. This 

staff is between 35 and 40 people and includes the shift supervisors in both operations and 

maintenance. The number of hourly personnel in mine operations could range from 90 to 110 

people during peak of operations. The number of mine maintenance personnel will range 

between 50 to 70 people depending on the maintenance philosophy adopted by the management 

(how much component replacement programs along with equipment dealer maintenance support 

is used). 

The Yerington District has a strong mining history and the state of Nevada has many active 

mines thus a skilled labor force is available for this operation. 

16.8 MINE CAPITAL COSTS 

The mine initial and sustaining costs are based on similar projects, estimated equipment 

requirements and file quotations for major equipment. Major ancillary equipment (dozers, 

graders, etc.), are shown as lot purchases as is support equipment (blasting truck, fuel trucks, 

pickups, cranes, etc.) and the engineering, geology and safety equipment (listed as other). The 

replacement of equipment is based on years of service versus hours of operation, which could 

change the replacement schedule. No mine capital purchases are made after year 11 and the last 

major equipment replacement is in year 8 when the initial fleet of 6 trucks, one drill and the 

loader are replaced. Table 16-9 shows the purchases of the mine capital equipment. Cost for the 

mine shops, warehouse and allowance for explosives storage are carried elsewhere in the overall 

capital estimate. 

Table 16-9: Mine Capital Estimate 

Purchased Units Pre-Prod 1 2 3 4 5 6 7 8 9 10 11 Total 

Drills 1 2 1 1 1 

Hyd Shovel 1 1 1 

Loader 1 1 

Trucks 6 1 1 3 1 6 1 

Major Aux Equip 1 1 0.5 

Mine Support Equip 1 1 

Other 1 

Capital Cost Unit Price 

$ x 1000 

Capital Purchases 

Drills 1,091 1,091 2,182 0 0 0 1,091 0 0 1,091 1,091 0 0 

Hyd Shovel 6,204 6,204 0 0 0 0 6,204 0 0 0 0 0 6,204 

Loader 2,900 2,900 0 0 0 0 0 0 0 2,900 0 0 0 

Trucks 2,564 15,384 0 2,564 2,564 0 7,692 0 2,564 15,384 0 0 2,564 

Major Aux Equip 16,200 16,200 0 0 0 0 0 0 16,200 0 0 8,100 0 

Mine Support Equip 5,000 5,000 0 0 0 0 0 5,000 0 0 0 0 0 

Other 1,000 1,000 0 0 0 0 0 0 0 0 0 0 0 

Total 47,779 2,182 2,564 2,564 0 14,987 5,000 18,764 19,375 1,091 8,100 8,768 131,174 

rounded 48,000 2,200 2,600 2,600 15,000 5,000 18,800 19,400 1,100 8,100 8,800 131,600 

16.9 MINE OPERATING COSTS 

An estimate of mine operating cost includes costs for drilling, blasting, loading and hauling plus 

ancillary activities (dump maintenance, road development and maintenance, pit clean up, etc.), 

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plus the mine services, mine maintenance and mine G&A departments. The general mine (mine 

services), general maintenance (maintenance supervision and maintenance of the smaller 

vehicles) and the mine G&A are estimated on a total cost per year basis with an increase in year 

6 as the mine begins the ramp up to the maximum material rate. The direct mining activities of 

drill ($0.15/t), blast ($0.20/t), load ($0.18/t), haul (range from $0.25/t to $0.47/t) and ancillary 

services ($0.22/t) are estimated on a cost per ton basis using information from similar size 

operations. The haul costs are escalated by year assuming longer hauls as the pits deepen and the 

waste dumps and heap leach facilities gain height. Table 16-10 is a summary of the mine 

operating costs by year. 

Table 16-10: Mine Operating Costs 

General General 

Year Total Drill Blast Load Haul 

Auxiliary Mine Maint. G&A Total $ Total $/t 

ktons 0.15 0.20 0.18 cost/ton $ x 1000 0.22 $ x 1000 $ x 1000 $ x 1000 $ x 1000 

Pre-Prod 500 75 100 90 0.25 125 1,000 500 500 1,000 3,390 6.78 

1 19,042 2,856 3,808 3,428 0.25 4,761 4,189 2,000 2,000 4,000 27,042 1.42 

2 17,174 2,576 3,435 3,091 0.27 4,637 3,778 2,000 2,000 4,000 25,517 1.49 

3 20,000 3,000 4,000 3,600 0.30 6,000 4,400 2,000 2,000 4,000 29,000 1.45 

4 20,000 3,000 4,000 3,600 0.32 6,400 4,400 2,000 2,000 4,000 29,400 1.47 

5 20,000 3,000 4,000 3,600 0.31 6,200 4,400 2,000 2,000 4,000 29,200 1.46 

6 30,000 4,500 6,000 5,400 0.32 9,600 6,600 2,500 2,500 4,000 41,100 1.37 

7 35,000 5,250 7,000 6,300 0.35 12,250 7,700 2,500 2,500 5,000 48,500 1.39 

8 35,000 5,250 7,000 6,300 0.37 12,950 7,700 2,500 2,500 5,000 49,200 1.41 

9 35,000 5,250 7,000 6,300 0.39 13,650 7,700 2,500 2,500 5,000 49,900 1.43 

10 35,000 5,250 7,000 6,300 0.33 11,550 7,700 2,500 2,500 5,000 47,800 1.37 

11 35,000 5,250 7,000 6,300 0.33 11,550 7,700 2,500 2,500 5,000 47,800 1.37 

12 32,403 4,860 6,481 5,833 0.35 11,341 7,129 2,500 2,500 5,000 45,644 1.41 

13 33,156 4,973 6,631 5,968 0.36 11,936 7,294 2,500 2,500 5,000 46,802 1.41 

14 35,000 5,250 7,000 6,300 0.37 12,950 7,700 2,500 2,500 5,000 49,200 1.41 

15 32,261 4,839 6,452 5,807 0.38 12,259 7,097 2,500 2,500 5,000 46,454 1.44 

16 27,749 4,162 5,550 4,995 0.40 11,100 6,105 2,500 2,500 5,000 41,912 1.51 

17 30,256 4,538 6,051 5,446 0.43 13,010 6,656 2,500 2,500 5,000 45,701 1.51 

18 22,612 3,392 4,522 4,070 0.45 10,175 4,975 2,500 2,500 5,000 37,134 1.64 

19 676 101 135 122 0.47 318 149 500 500 500 2,325 3.44 

Total 515,829 77,372 103,165 92,850 0.354 182,762 114,372 43,500 43,500 85,500 743,021 1.44 

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17 RECOVERY METHODS 

17.1 OVERVIEW OF PLANNED FACILITIES 

The process facilities planned for the MacArthur Copper Project include a ROM heap leach 

facility to recover copper in a leach solution, and a solvent extraction and electrowinning 

(SX/EW) facility to recover the copper from the leach solution and produce a cathode quality 

copper for sale. Also included is a sulfur burning sulfuric acid plant with a power plant to 

generate electrical power from the waste heat produced from the combustion of sulfur. Other 

facilities include solution ponds, water and power distribution, and infrastructure to support the 

facilities. An overall flow sheet for the heap leach and SX/EW facilities is shown in Figure 17-1 

at the end of this section. 

17.2 HEAP LEACH PAD 

MacArthur ore will be mined from open benches, loaded into mine haul trucks, transported 

directly to the heap leach pad and stacked. The ore will be dumped on the leach pad and irrigated 

with an acidified leach solution (raffinate). Raffinate will be pumped from the raffinate pond 

through a pipeline and distribution network to drip emitters which will distribute the leach 

solution to the surface of the ore pile on the leach pad to minimize evaporation losses. Some 

sprays may be used on side slopes or to increase evaporation if required to maintain the process 

water balance. The leach solution will percolate through the ore pile and dissolve soluble copper 

from the ore before being directed along the impermeable leach pad liner system to the solution 

collection system. 

The heap leach pad design will conform to the Nevada Division of Environmental Protection 

(NDEP) requirements and will consist of, from bottom to top, a compacted soil base, covered by 

clay or a geosynthetic clay liner (GCL), an HDPE liner, and HDPE perforated piping network of 

collection pipes with 24 inches of crushed over-liner to protect the collection piping. 

Copper bearing leach solution, called pregnant leach solution (PLS), will flow by gravity from 

the leach pad collection system to a lined collection pond (pregnant pond). The PLS will be 

pumped at a rate of 10,400 gallons per minute with 1 gpl copper to the solvent extraction mixersettlers. 

The pump discharge pipes will be combined in a single pipeline to the solvent extraction 

circuit. 

While one lot of ore is being leached, the next will be mined and placed in another section of the 

leach pad. When an ore lot has completed the primary leach cycle, solution application will be 

transferred to the next lot of ore. When all the ore on a lift (or layer) has been leached, additional 

ore lifts will be placed on top of the previous lift and leaching will continue. The process of 

layering and leaching the ore will be repeated to a maximum acceptable number of ore lifts on 

the leach pad. If required, one inner-lift liner may be used at one-half the pad height. 

A storm water pond will be installed to handle excess water that might occur during a large 

precipitation event. The PLS collection pond will be designed to overflow to the storm water 

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pond which is sized to accommodate a 100 year storm event. Water that may accumulate in the 

storm water pond will be periodically pumped to the raffinate solution pond. 

17.3 SOLVENT EXTRACTION 

Copper contained in the aqueous phase PLS will be extracted by contact with organic reagents 

carried in an organic solution (organic phase) in the solvent extraction circuit. Copper transferred 

to the organic phase will be stripped from the organic solution by contact with an acidic 

electrolyte solution (lean electrolyte) that will have circulated through the electrowinning cells. 

This transfer of copper enriches the electrolyte solution to form the rich electrolyte. The rich 

electrolyte will be pumped to the electrowinning cells for copper electrowinning onto stainless 

steel cathode blanks. Copper loaded on the stainless steel blanks will be harvested from the 

electrowinning cells on a weekly schedule. Copper will be removed from the stainless steel 

blanks by a stripping machine. Copper plates produced by this process, LME Grade A, will be 

weighed and bundled into 2 to 3 ton packages for shipment to market. 

The solvent extraction plant will consist of one train of mixer-settlers. The train will have three 

stages of extraction arranged in a series parallel configuration and one stage of stripping. 

Aqueous and organic streams will flow counter-currently in extraction. 

The PLS will be divided into two streams. One stream enters the two stage extraction pumper 

mixers, operated in series, and is contacted with stripped organic. After the two phases, organic 

and aqueous, have been mixed by flowing through mix tanks in series, the resulting mixture will 

be discharged into a single extraction settler to allow the two phases to disengage. The organic 

solution will float on top of the aqueous solution (raffinate) allowing the two phases to be 

separated by a weir system at the discharge end of the settler. The partially loaded organic then 

passes on to two other mixer-settlers, operated in series, where it counter-currently contacts the 

other half of the remaining leach solution to produce another raffinate stream and loaded organic. 

The resulting copper depleted aqueous solution, or raffinate, will flow by gravity to the raffinate 

pond. 

The stripping mixer-settler will process loaded organic solution to remove copper extracted from 

the aqueous solution. Loaded organic solution from the first stage of extraction will flow by 

gravity to a loaded organic tank. Loaded organic solution will be pumped to the stripping stage 

pumper mixer and will be mixed with lean electrolyte solution from electrowinning. The same 

flow pattern occurs in the stripping circuit as in each stage of the extraction circuit. Organic 

solution will enter the strip stage mix tank and will be mixed with lean electrolyte solution. After 

the two phases, organic and aqueous will have been mixed by flowing through mix tanks in 

series, the resulting mixture will be discharged into settlers to allow the two phases to disengage. 

The organic solution will float on top of the aqueous solution allowing the two phases to be 

separated by a weir system at the discharge end of the settler. The stripped organic solution will 

flow to the second stage of extraction. The aqueous enriched electrolyte solution will be split 

with a portion advancing to electrowinning and the balance recycled within the organic strip 

stage. 

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17.4 ELECTROWINNING 

The rich electrolyte solution from solvent extraction will flow by gravity to the electrolyte filter 

feed tank and will be pumped through two electrolyte filters operating in parallel. The filters will 

be backwashed periodically with lean electrolyte solution and air from a scour air blower. Filter 

backwash solution will be returned to the extraction settlers. 

Filtered electrolyte solution will be pumped from the filtered electrolyte storage tank to the 

electrolyte heating circuit. The filtered electrolyte will flow through two heat exchangers 

operating in series. In the first heat exchanger, electrolyte will be warmed by lean electrolyte 

returning to solvent extraction from electrowinning. In the second heat exchanger, electrolyte 

will be heated, if necessary, with supplemental heat to final temperature for electrowinning. A 

hot water heating system will be installed to provide supplemental heat. 

A diesel fired steam boiler will heat water in a hot water tank through a steam loop from the 

boiler. Hot water will be circulated through the heat exchanger when additional heat is required 

to heat the electrolyte solution, as during start-ups. 

The electrowinning circuit tank house will contain 54 electrowinning cells. Heated electrolyte 

solution will enter the electrowinning cell circuit by flowing to an electrolyte recirculation tank 

where it will be mixed with electrolyte solution returning from the other electrowinning cells. 

The electrolyte recirculation tank will be a two compartment tank. The lean electrolyte from 

electrowinning will return to the smaller compartment that contains a pump connection for 

returning the electrolyte to the stripping circuit. The excess solution will overflow a baffle and be 

mixed with rich electrolyte and will be pumped to electrowinning cells. Lean electrolyte will be 

pumped from the recirculation tank through the heat exchanger and to the strip stage in the 

solvent extraction circuit. 

Copper will be plated onto stainless steel cathode blanks. A cathode stripping machine will be 

used to remove the copper plates from the stainless steel blanks. The cathode stripping machine 

will perform several steps in sequence; cathode washing, hammering and flexing the blanks to 

loosen the copper plates, stripping the plates from the blanks, stacking and banding the copper 

plates, and stainless steel blank preparation for return to the electrowinning cells. 

The tank house will have an overhead bridge type crane for transporting cathodes and anodes to 

and from the cells. 

A filter system will be installed to process solvent extraction “crud” (the material that forms 

from an accumulation of solids and organic and aqueous solution at the organic/aqueous 

interface in the settlers) to separate organic solution that will be reused in solvent extraction. 

Crud will be drained or decanted from the settlers to a crud holding tank. When sufficient 

material has been collected it will be pumped to a crud decant tank. The crud will be diluted with 

diluent and alternatively mixed and allowed to settle in the decant tank. The organic layer that 

will form in the decant tank will be pumped from the tank to the loaded organic tank. Sediment 

from the crud holding tank will be pumped to a mix tank where it will be mixed with 

diatomaceous earth filter media. The mixture will be pumped to a plate and frame filter. Filtrate 

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will be collected in a tank. Filtrate will be periodically transferred, depending on the phase 

content of the filtrate, to the solvent extraction aqueous system or to the solvent extraction 

organic system. 

17.5 SULFURIC ACID PLANT 

Molten sulfur will be received at a rail siding offsite in rail tank cars of approximately 100 ton 

capacity. The rail cars must be heated by steam to liquefy the sulfur since heat loss in the car 

during transit will solidify some of the sulfur. When re-heated, the molten sulfur is discharged to 

a receiving pit and pumped into a heated storage tank. The molten sulfur will be transferred from 

the heated storage tank at the rail siding by truck and tanker to heated storage tanks at the 

sulfuric acid plant. 

Molten sulfur is pumped from the storage tanks at the acid plant to the sulfur furnace where it is 

mixed with high pressure air to atomize the sulfur and air to combust the sulfur. A bleed stream 

of sulfur recirculates back to the sulfur storage tanks to ensure a consistent feed of sulfur to the 

sulfur burners. Excess air is provided at the burners to ensure complete combustion and 

sufficient excess oxygen in the off gas for the conversion of SO2 to SO3 in the acid plant. The 

combustion process in the sulfur burner produces an off-gas at about 11% SO2. 

The combustion air for the sulfur furnace is first dried to remove any moisture in the air prior to 

combustion. This is to prevent corrosion in the rest of the downstream equipment. Ambient air is 

drawn into an air inlet filter and silencer ahead of the main acid plant blower and then delivered 

to the bottom of a packed drying tower by the main acid plant blower. In the drying tower, the 

ambient air flows through a packed section of ceramic saddles in countercurrent flow with 96% 

H2SO4. The air leaves the top of the drying tower and is delivered to the sulfur furnace for 

combustion. The circulating acid, at 96% H2SO4, will absorb the water in the air, which will 

tend to reduce the acid strength in the drying tower. This will be offset by a cross bleed of higher 

strength acid from the absorption towers downstream. Excess 96% H2SO4 generated in the 

drying tower is advanced to the absorption towers. 

Off gas from the combustion process in the sulfur furnace will pass through a fire tube waste 

heat boiler and then to the converter. Steam is generated in the waste heat boiler which can be 

used to generate electrical power, discussed later. The converter is a four bed converter with 

vanadium pentoxide catalyst in each bed. As the gas passes through the catalyst beds, SO2 is 

converted to SO3. The reaction is exothermic and increases the temperature of the gas. After 

each pass through a converter bed, the gas is cooled through gas-to-gas heat exchangers to cool 

the gas for the next pass. After the second or third pass, approximately 85% of the SO2 has been 

converted to SO3 and the gas then passes to an intermediate absorption tower before returning to 

the final one or two passes. At the outlet of the fourth pass, the gas then passes to the final 

absorption tower. 

The intermediate and final absorption towers are similar in design as the drying tower. The gas 

stream from the converter enters the absorption towers at the bottom and passes through a 

section of ceramic packing saddles where the SO3 comes in contact with, and absorbed by, the 

circulating 98.5% H2SO4. The gas leaves the absorption tower at the top and returns to the 

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converter for further SO2 conversion (in the case of the intermediate absorption tower) or is 

discharged to atmosphere through a stack (in the case of the final absorption tower). As the SO3 

is absorbed into the 98.5% circulating acid, the acid will tend to gain in strength. A cross bleed 

of the higher strength acid from the absorption towers is directed to the drying tower to maintain 

96% acid strength in the drying tower. Excess 96% acid in the drying tower advances to the 

intermediate and final absorption towers. Water is also added to the absorption towers to 

maintain a constant acid strength of 98.5% acid. Excess 98.5% acid in the absorption towers is 

pumped to storage as the final product from the acid plant. 

The double-contact double-absorption sulfuric acid plant conforms to the Best Available 

Demonstrated Control Technology (BADCT) for controlling sulfur dioxide emissions. 

17.6 POWER PLANT 

The power plant will be located adjacent to the sulfuric acid plant. Steam from the waste heat 

boiler will drive a steam turbine generator to generate electrical power. The turbine exhaust will 

be directed to a shell and tube steam condenser operating under vacuum. The condensate from 

the steam condenser is collected and pumped through a water treatment system to maintain boiler 

quality water. A dump condenser will also be provided to condense steam in the event the 

turbine generator cannot accept the steam from the waste heat boiler. Boiler blow down will be 

cooled and directed to the raffinate pond. 

The power plant turbine generator will be connected to the main electrical substation buss for 

distribution to the SX/EW and acid plant facilities. 

Fresh water make-up to the steam system will be treated by filtration to remove particulates, 

cation exchange water softening to remove scale producing ions, chemical treatment of the 

softened water with a dispersant and anti-scalant, and Reverse-Osmosis (RO) filter to achieve the 

desired water quality. The circulating condensate water will be treated through a condensate 

polishing system as a precautionary measure to ensure boiler quality water is maintained. 

Cooling towers will be required to provide cooling water for the sulfuric acid plant acid coolers 

and the dump condensers at the power plant. The cooling towers will have chemical water 

treatment for the fresh water make-up. Fresh water will be required for make-up to the system to 

account for the evaporation loss and blow down. Blow down from the cooling towers is required 

to maintain a proper level of dissolved solids in the cooling towers. The blow down will be 

directed to the raffinate pond. 

17.7 ANCILLARY FACILITIES 

Ancillary facilities to support the MacArthur process facilities include fuel storage and 

distribution systems for heavy equipment and light vehicles, an electrical substation, a 

guardhouse and truck scale at the entrance to the property. Existing buildings at the Yerington 

site will be refurbished and used for administration, an analytical lab, a mine truck shop, 

warehouse, and maintenance facilities. 

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Figure 17-1: Overall Process Flowsheet 

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18 PROJECT INFRASTRUCTURE 

18.1 SITE LOCATION 

The MacArthur Copper Project is located near the geographic center of Lyon County, Nevada, 

along the northeastern flank of the Singatse Range, approximately seven miles northwest of the 

town of Yerington, Nevada. The property is accessible from Yerington by approximately five 

miles of paved roads and two miles of Lyon-county maintained gravel road. The nearest major 

city is Reno, Nevada, approximately 75 miles to the northwest. 

The process facilities are located east of the ore body and west of Alternate Highway 95. The 

heap leach pad is directly east of the MacArthur pit with the process facilities located south of 

the heap leach pad. The area covered by the heap leach pad is approximately 390 acres and the 

process facilities occupy another 85 acres. The heap leach pad and process facilities are shown in 

Figure 18-1 at the end of this section. 

18.2 PROCESS BUILDINGS 

The process facility generally consists of solvent extraction settlers, a tank farm, and an 

electrowinning building. The solvent extraction settlers are four covered tanks approximately 60 

feet by 115 feet and 4 foot deep. The tank farm is located below the solvent extraction facilities 

and contains all the circulation tanks, pumps, heat exchangers, and filters that service the solvent 

extraction and electrowinning facilities. The electrowinning building is a pre-engineered steel 

building with corrugated metal roofing and siding. The main cell area is approximately 150 feet 

long and 70 feet wide holding two rows of 27 electrowinning cells each. A building extension, 

approximately 98 feet by 180 feet, is located at one end of the electrowinning cells to house the 

automatic stripping machine and the cathode handling equipment. An overhead crane in the 

building services the electrowinning cells and stripping machine. An electrical equipment room 

and control room is located on one side of the cathode handling section which overlooks the 

cathode stripping operation. Cathode handling, weighing and banding is performed at the other 

side of the cathode handling section. An asphalt paved laydown area is provided outside the 

cathode handling area to allow cathode storage and loading of cathodes onto flatbed trailers for 

shipment to market. The building is provided with ventilation fans and scrubbers to ventilate the 

cell area. 

The control room will have space for offices, a laboratory, and restrooms. The laboratory will be 

for analyzing routine shift samples. An existing building in the Yerington operation will be 

refurbished for a main analytical laboratory to supplement the laboratory at in the tank house. 

The grey and black water from the restrooms will report to a dedicated septic system. 

18.3 ANCILLARY BUILDINGS 

Ancillary buildings necessary to support the MacArthur Copper Project include an 

administration building, a warehouse / maintenance building, an analytical laboratory, mine truck 

shop, a change house, fuel storage and dispensing facilities, and the main gatehouse with truck 

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scale. Some ancillary buildings at the existing Yerington facility will be refurbished and used to 

support the MacArthur Copper Project. 

18.3.1 Administration Building 

Quaterra occupies offices in the town of Yerington that will continue to be used for some of the 

administrative functions. A smaller building at the Yerington mine site will be re-furbished to 

provide additional offices for onsite supervisors. An allowance was provided to include 

approximately 1,200 square feet of office space at the mine site. 

18.3.2 Warehouse / Plant Maintenance Building 

An existing industrial building at the Yerington mine site will be converted to a maintenance 

building and warehouse. The building is approximately 12,000 square feet and will be 

partitioned in the center to provide 6,000 square feet for warehouse space and 6,000 square feet 

for maintenance space. The building is of steel construction and corrugated roofing and siding. 

Metal shelving will be provided for the warehouse and the maintenance side will have offices 

and restrooms and will house the plant maintenance facilities. A fenced area will be provided 

outside the warehouse for secure outdoor storage. 

18.3.3 Analytical Laboratory 

An analytical laboratory will be provided at the Yerington mine site to supplement the SX/EW 

laboratory. An existing building, approximately 1,500 square feet, will be re-furbished and 

provisioned with sample preparation equipment, laboratory equipment, ventilation systems and 

offices. Mine samples will be processed at this laboratory. 

18.3.4 Mine Truck Shop 

An existing mine truck shop exists at the Yerington mine site and can continue to be used. The 

building is approximately 9,000 square feet and is equipped with an overhead crane. An 

allowance was provided for re-furbishing of the building and an allowance for new equipment. 

18.3.5 Change House 

A new change house building will be provided at the MacArthur SX/EW facility. The change 

house is a pre-engineered steel building with corrugated roofing and siding. The building is 

approximately 1,000 square feet. 

18.3.6 Main Gatehouse 

A new modular building will be provided at the main gate to control access to the SX/EW plant. 

The building is 44 feet by 14 feet with 10 foot eaves. A truck scale will be provided at the main 

plant entrance to weigh all receipts of reagents and consumables as well as cathode copper 

production leaving the plant. 

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18.3.7 Fuel Storage and Dispensing 

Fuel storage and dispensing facilities will be provided for mine trucks and in-plant vehicles. 

Two 50,000 gallon diesel storage tanks are provided to service the mine trucks and mining 

equipment. Two 5,000 gallon storage tanks will also be provided for small vehicles and 

equipment. One tank will be for diesel and the second tank for gasoline. Fuel will be received by 

tank trucks from Reno, Nevada or other nearby source and dispensed at site. 

18.4 ACCESS ROADS 

Access to site is from Alternate Highway 95, west on Luzier Lane, and north on Mason Pass 

Road. Entrance to the SX/EW facility is off of Mason Pass Road. Once on site existing haul 

roads connect to the mine site and the existing Yerington facilities. No new access roads will be 

required. 

18.5 RAILROAD FACILITIES 

A railroad and siding exists at Wabuska, Nevada, approximately 10 miles north of Yerington on 

Alternate Highway 95. The rail line connects from Hawthorn, Nevada to Salt Lake City, Utah, 

generally following Interstate Highway 80 east. The existing siding will be used to receive 

molten sulfur by rail, with the sulfur transferred from rail cars to truck and tankers at Wabuska 

for the final transport to site. Molten sulfur will be received in 100 ton rail cars at the rate of 

approximately 14 rail cars per week. 

18.6 POWER SUPPLY & DISTRIBUTION 

Power for the facility will be taken from an existing 69 kV power line feeding from the existing 

Fort Churchill generating facility to the town of Yerington, a distance of approximately 10.5 

miles. The 69kV power line approaches the SX/EW plant site along the eastern project boundary. 

A tap will be taken from the existing power line and a short, 800-foot long, power line will be 

constructed to connect to the SX/EW main electrical substation. The condition of the existing 69 

kV power line from Fort Churchill has not been assessed at this time and may require upgrades 

in order to service the MacArthur Project site. 

At the SX/EW main substation, power will be transformed to 34.5 kV or 13.8kV for distribution 

throughout the plant. Additional transformers will be provided in the various process areas to 

provide medium voltage (4160 V) and low voltage (480 V) to feed the end users. Electrical 

equipment rooms and motor control centers will be located at solvent extraction, tank farm, 

electrowinning, acid storage, and the solution ponds. Step down transformers will be provided to 

serve the change house, guard house and fuel storage facilities. 

18.7 WATER SUPPLY & DISTRIBUTION 

The total water consumption for the project is estimated to be approximately 1,600 gpm, or 2,590 

ac-ft per year. Quaterra and its subsidiary companies own approximately 8,600 ac-ft of water 

rights at the Yerington Mine Site. Fresh water for the project will be taken from wells on or near 

the MacArthur property and pumped to a 380,000 gallon fresh water/fire water storage tank. The 

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lower 120,000 gallons in the storage tank will be reserved for fire water. Fresh water for plant 

use will be taken from the storage tank above this reserve level for fire suppression. An 

additional 10,000 gallon potable water tank will be provided to service the potable water system. 

The fresh water will be filtered and chlorinated before stored in the potable water tank and 

distributed throughout the plant. Potable water will be used for offices, labs, restrooms, and eye 

wash stations. Fresh water will be used for fire suppression, wash down, and process water 

make-up. The facility will be designed as a zero discharge facility. 

18.8 WASTE MANAGEMENT 

It is assumed a private landfill will be provided on the property for non-hazardous solid waste. 

This facility will not accept any off site wastes and will be used primarily for construction debris, 

non-putrescible materials and waste from maintenance and operations meeting the definition of 

inert or non-hazardous materials; such as air filters, gloves, boxes, non-recyclable packaging 

material, hoses, piping, etc. 

Recyclable materials that are non-hazardous, such as scrap metal, paper, used oil, batteries, wood 

products, etc., will be collected in suitable containers and disposed of through recyclers. 

Hazardous materials such as contaminated greases, chemicals, paint, reagents, etc. will be 

collected shipped off-site for destruction or disposal. Some hazardous materials, such as lead 

flakes and anodes, may also be recycled through appropriate recyclers. 

18.9 SURFACE WATER CONTROL 

Storm water run-off will be diverted around the plant facilities as much as possible. The natural 

gradient of the heap leach area generally slopes to the northeast. The SX/EW facilities slope to 

the south east. A diversion channel is provided to direct non-impacted run-off water around or 

through the plant without contacting impacted areas. Impacted run-off water from within the 

plant will flow to the tank farm area and collected in the tank farm sump. The impacted water 

will be pumped from the tank farm sump to the raffinate pond. A storm water pond is located 

adjacent to the PLS pond to accept any overflow from the PLS pond during storm events. The 

overflow will then be pumped back to the process or the raffinate pond. Annual precipitation 

ranges from 5 to 8 inches per year, more in the higher elevations. 

18.10 TRANSPORTATION & SHIPPING 

All materials coming into the plant will be by truck, including the molten sulfur transferred from 

the rail facilities at Wabuska. The facility will also be able to receive sulfuric acid by truck in 

emergencies, if needed, with the acid unloaded into the same sulfuric acid storage tanks. 

Incoming materials include reagents, extractant, kerosene, gasoline, diesel, warehouse stock and 

spare parts. 

The primary product leaving the plant is cathode copper, which will be by flatbed tractor trailers. 

Recycle materials leaving the plant will also be by truck. Scales to weigh full loads and empty 

loads into and out of the plant are provided at the main gate for the highway trucks. The 

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approximate quantity of major products and consumables entering and leaving the plant are 

shown in Table 18-1 below. 

Table 18-1: Products & Consumables 

Quantity 

Quantity 

Material 

per Year Units per day Units 

Extractant 38,700 lbs. / yr. 106 lbs. / day 

Diluent (Kerosene) 516,500 lbs. / yr. 1,415 lbs. / day 

Cobalt Sulfate 59,500 lbs. / yr. 163 lbs. / day 

Guar 107,700 lbs. / yr. 295 lbs. / day 

Sulfur for Sulfuric Acid 77,400 tons / yr. 212 tons / day 

Cathode Copper (Production) 20,700 ton / yr. 57 tons / day 

18.11 COMMUNICATIONS 

The connection to telephone and internet services for the project has not been confirmed at this 

time; however, telephone service is available at the town of Yerington and the existing Yerington 

mine site. It is assumed that the telecommunication system will be integrated with the onsite data 

network system utilizing a voice over I/P (VoIP) phone system. A dedicated server will be 

provided for setup and maintenance of the VoIP system and for accounting of all long distance 

phone calls. Handsets will plug into any network connection in the system for 

telecommunications. The office Ethernet network will support accounting, payroll, maintenance 

and other servers as well as individual user computers. High bandwidth routers and switches 

will be used to logically segment the system and provide the ability to monitor and control traffic 

over the network. 

A process control system Ethernet network will support the screen, historian and alarm servers 

connected to the control room computers as well as Programmable Logic Controllers (PLC). 

This system will incorporate redundancy and a gateway between the office system and control 

system to allow business accounting systems to retrieve production data from the control system. 

No phone or user computer will be connected to this system. 

The internal communications within the plant will utilize the same VoIP phone system, which 

will provide direct dial to other phones throughout the plant site. Mobile radios and cell phones 

will also be used by operating and maintenance personnel for daily communications while 

outside the office. 

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Figure 18-1: MacArthur Heap Leach and Process Facilities 

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19 MARKET STUDIES AND CONTRACTS 

Copper is an international traded commodity with the price governed by the worldwide balance 

of supply and demand. The copper price is determined by the major metals exchanges; consisting 

of the New York Mercantile Exchange (COMEX), the London Metals Exchange (LME), and the 

Shanghai Future Exchange (SHFE). Recent historical copper prices are shown in below. 

Figure 19-1: Historic Copper Price 

The final product from the MacArthur facilities will be high purity (99.9%) electrolytic cathode 

copper in sheets of about 100 pounds each; bundled into approximately 55 cathode sheets or 

5,500 pounds per bundle. Approximately 90% of copper cathode production in the United States 

goes to wire rod mills and eventual wire production and to brass mills producing various copper 

and copper alloy shapes. North America is a net importer of refined copper with a projected 

consumption of refined copper for 2012 of 2.5 million tons and a production of 2.0 million tons. 

The 0.5 million ton shortfall is made up of imports primarily from Chile, Canada, Peru, and 

Mexico. It is expected that the production from the MacArthur facilities would be absorbed into 

the North American copper market for refined copper, displacing the import copper. 

Typical terms related to cathode copper shipping include FCA (Free Carrier) at the refinery; that 

is the buyer arranges and pays for cathode transportation from the refinery and the seller loads 

the cathode onto buyer’s trucks. The price is based on the average COMEX price during the 

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Quotation Period plus a premium for ASTM Grade 1 quality material, with negotiated discounts 

for lesser quality material. The net premium is the quoted premium, less freight charges and a 

margin allowed to the merchant buyer. The Quotation Period is the month of shipment or the 

month following the month of shipment. Payment is typically 2 days after the date of shipment. 

Northeastern Texas is the major regional market for western US cathodes. 

Transportation is by flatbed truck for shorter distances or by rail in box cars for the longer 

distances. 

A formal market study has not been conducted in this phase of the project and there are no 

established contracts for the sale of copper cathode in place at this time. 

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20 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR 

COMMUNITY IMPACT 

20.1 ENVIRONMENTAL LIABILITIES 

Previous mining at the MacArthur site was conducted by Arimetco in the late 1990’s, and 

included the construction of open pits, a waste rock dump, and access roads. The waste rock 

dump was reclaimed by the U.S. Bureau of Land Management following Arimetco’s bankruptcy 

in 1997 and abandonment of the site in 2000. Numerous historic adits and underground workings 

are located throughout the project area, many of which have been secured by the Nevada 

Division of Minerals (in coordination with Quaterra) to prevent unauthorized access. 

Extensive exploration has also occurred throughout the project area since the 1970’s. Exploration 

related disturbance, including both historic disturbance and new disturbance created by Quaterra, 

consists of historic drill sites, trenches, and numerous drill roads. Quaterra has a reclamation 

bond that covers exploration and is responsible only for the exploration disturbance created 

during their tenure at the site since 2007. 

The ore from the previous MacArthur Pit was processed through heap leaching at the Yerington 

Mine facility which is located approximately five miles from the project area. These materials 

are not associated with the current operation at MacArthur and are being evaluated as a potential 

resource for reprocessing at the Yerington site (see Section 24). 

Because the site is basically at or near elevation, no pit lake formed following the cessation of 

mining by Arimetco. Although uncertain, it is unlikely that a pit lake would form as a result of 

Quaterra’s proposed mining operations at MacArthur. However, additional hydrogeological 

investigations will be necessary before a final determination can be made in this regard. 

There are no known liabilities to which the MacArthur property is subject. 

20.2 PERMITS 

The mineral resources at MacArthur are located on public (unpatented) mining claims 

administered by the U.S. Department of the Interior, Bureau of Land Management, Carson City 

District, Sierra Front Field Office (BLM). Mine development will require participation of the 

BLM as the primary land manager. Various departments within the State of Nevada will be 

cooperating agencies in permitting mining development and process facilities at the site. Based 

on the current mine plan, and proposed facilities, the following Table 20-1 lists the principal 

permits necessary to commence mining operations. To date, none of these permits have been 

acquired for mining operations, although permitting for exploration activities is complete. 

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Table 20-1: Summary of Major Permits for Future Mining 

Regulatory Agency Permit Name 

Bureau of Land Management 

Bureau of Alcohol, Tobacco, Firearms, and 

Explosives 

Mine Safety and Health Administration 

Environmental Protection Agency 

Federal Permits 

• Approved Plan of Operations/Decision Record 

• Roads and utility Rights-of-Way 

• Authorization to purchase, transport, or store 

explosives 

• Notification of Commencement of Operation 

• Employee and Facility Health and Safety 

• Hazardous Waste ID No. (small quantity 

generator) 

State Permits 

Nevada Division of Environmental Protection 

Bureau of Mining Regulation and Reclamation 

• 

• 

Water Pollution Control Permit 

Reclamation Permit 

• Class I (PSD) or Class II Permits to Construct 

Bureau of Air Pollution Control 

and Operate 

• Mercury Permit 

Bureau of Water Pollution Control 

• 

• 

Stormwater NPDES General Permit 

Septic Permit 

• Approval to Operate a Solid Waste System (if 

Bureau of Waste Management 

necessary) 

• Hazardous Waste Management Permit 

Bureau of Safe Drinking Water • Potable Water Permit 

Nevada Division of Water Resources 

• Permit to Appropriate Water 

• Permit to Construct a Dam 

• Mineral Exploration Hole Plugging 

Nevada Department of Wildlife 

• Industrial Artificial Pond Permit 

State Fire Marshall 

• Hazardous Materials Permit 

Local Permits 

Lyon County 

• Special Use Permit 

• Building Permit 

• Business License 

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20.2.1 Federal Permitting 

A mine plan of operations (PoO) will be prepared to describe the construction, operation, and 

reclamation of each facility along with a cost estimate that presents the reclamation and closure 

costs if the BLM were required to take over reclamation of the mine site. Information required 

for the PoO includes: pit location(s) and lateral and vertical extent of disturbances; heap leach 

pad conceptual designs; location of haul roads, ore stockpiles, waste rock dumps, growth media 

stockpiles, office/laboratory, shops, diesel/lubricant storage and distribution system, well and 

associated piping; power line locations; generators; schedule of construction and operation; 

mining schedule; and equipment list. A reclamation plan is an important part of the PoO, which 

describes the activities that will take place to estimate the reclamation cost for bonding. The PoO 

will also function as the Reclamation Permit application for the State of Nevada (BMRR) (see 

below). 

The PoO provides sufficient detail to identify and disclose potential environmental issues during 

the National Environmental Policy Act (NEPA) review, including an environmental impact 

statement (EIS) or Environmental Assessment (EA). The BLM will likely require an EIS for a 

project of this type. As a general rule, the PoO/EIS process for a mining/mineral beneficiation 

project is a minimum 36 months process (including 12 months for baseline data collection and 

PoO development plus a minimum of 24 months on review and EIS preparation). However, 

internal BLM situations could occur beyond the control of the project proponent, and a number 

of potential external events (public or cooperating agency opposition) could lengthen the overall 

EIS schedule. It is not uncommon for a mining PoO/EIS process to be three to five years before a 

Record of Decision (ROD) is issued. 

The BLM has recently implemented new procedures requiring that at least one year of baseline 

data be submitted with the PoO in accordance with the state-wide Instruction Memorandum No. 

NV-2011-004 (dated November 5, 2010). The purpose of this guidance is to “improve the 

efficiency and effectiveness of processing mine Plans of Operation.” To that end, the BLM front 

loaded the permitting process for the collection of baseline data and environmental studies before 

the PoO is submitted for BLM review and NEPA analysis. BLM believes this should reduce the 

review period and overall NEPA process. 

The requirements of the BLM PoO document are fairly well defined. However, baseline data 

necessary for the impact assessment phase of the project will need to be collected, analyzed, and 

interpreted in conjunction with the BLM to ensure the information collected meets the Data 

Quality Objectives (DQOs) of the program. Longer-lead items to be considered include: 

• Groundwater sampling (hydrogeology) in the project area for depth and quality (for use 

in both the NEPA analysis and the State’s Water Pollution Control Permit application); 

and 

• Geochemical characterization of waste rock, ore, and spent leach material including acidbase 

accounting (ABA), meteoric water mobility procedures (MWMP) testing, and 

humidity cell (HCT) testing. The geochemical characterization program must be 

approved in advance by the BLM and the NDEP, and be in accordance with BLM 

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Instruction Memorandum (IM) No. NV-2010-014 Nevada Bureau of Land Management 

Rock Characterization Resources and Water Analysis Guidance for Mining Activities 

(January 8, 2010). 

The collection of environmental baseline data necessary for development of the mine operations 

PoO and EIS review process was initiated with an expanded vegetation monitoring program in 

May 2012. Vegetation, including special status species, is a time-critical environmental element 

with limited windows for data collection. It is generally the first element initiated as part of a 

baseline data collection program. Other “elements of the environment” (BLM National 

Environmental Policy Act Handbook H-1790-1, 2008) to be considered during the NEPA 

process include: 

• Air Quality, 

• Areas of Critical Environmental Concern, 

• Cultural Resources, 

• Environmental Justice, 

• Floodplains, 

• Grazing Management, 

• Land Use Authorization, 

• Migratory Birds, 

• Minerals, 

• Native American Religious Concerns, 

• Noxious Weeds, Invasive and Non-Native Species, 

• Paleontological Resources, 

• Recreation, 

• Social and Economic Values, 

• Soils, 

• Special Status Species (plants and animals), 

• Threatened and Endangered Species (plants and animals), 

• Vegetation, 

• Visual Resources, 

• Wastes (solid and hazardous), 

• Water quality (surface and ground), 

• Wetlands/Riparian Zones, 

• Wild Horses and Burros, 

• Wilderness and wilderness characteristics, and 

• Wildlife. 

Table 20-2 presents a list of elements or studies that generally require more detailed 

investigations and may need to be undertaken during the mine planning phase in advance of the 

NEPA process. These studies will also be used to support the acquisition of various other 

operating permits. Many of these studies were performed within the project site as part of the 

PoO/EA for Exploration and may have to be updated for the EIS (BLM, 2009). 

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Water 

Geology and 

Geochemistry 

Cultural 

Resources 

Biological 

Resources 

Table 20-2: Future Baseline Studies 

Permit/Authorization Investigations/Studies 

• NEPA Analysis 

• Water pollution control 

permit 

• Stormwater control 


• Water pollution control 

permit 

• Waste rock dump design 

• Dump and heap closure 

• Closure planning for dumps, 

heaps, and tailings 



Monitor surface waters in project vicinity on a 

seasonal basis for quality and quantity 

Monitor groundwater for level and water quality 

especially in the pit, dump, and heap areas to 

collect baseline quality data 

Collect representative samples of waste rock, 

ore, and spent heap ore for geochemical 

characterization (ABA, MWMP, and HCT) 

Condemnation drilling in proposed locations of 

facilities 

Conduct a Class III survey in previously 

unsurveyed or as directed by the BLM 

Mitigate sites that cannot be avoided 

Determine presence or absence of threatened, 

endangered, or special status plant and animal 

species including golden eagles in previously 

unsurveyed areas 

Determine presence or absence of game species 

Other federal permits that may be required include a hazardous waste identification number from 

the U.S. Environmental Protection Agency and an explosives use permit from the Bureau of 

Alcohol, Tobacco, Firearms, and Explosives. 

20.2.2 State Permitting 

The State of Nevada requires permits for all mineral exploration and mining operations 

regardless of the land status of the project. The two most important operational permits include 

the Water Pollution Control Permit (WPCP) and the Reclamation Permit; both issued by the 

Department of Conservation and Natural Resources, Division of Environmental Protection, 

Bureau of Mining Regulation and Reclamation (BMRR). The BMRR is composed of three 

distinct technical branches; Regulation, Closure, and Reclamation, and its mission is to ensure 

that Nevada's waters are not degraded by mining operations and that the lands disturbed by 

mining operations are reclaimed to safe and stable conditions to ensure a productive post-mining 

land use. 

The Regulation Branch of the BMRR issues a WPCP to a mine operator prior to the construction 

of mining, milling or other beneficiation processes. Facilities utilizing chemicals for processing 

ores are generally required to meet a zero discharge performance standard to protect of surface 

waters, which standard requires containment of all process fluids. The WPCP covers mine 

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facility components including buildings, structures, facilities or other installations from which 

there is or may be a discharge of pollutants. In the case of the MacArthur Copper Project, this 

includes, to the following facilities: 

• Acid leach pad; 

• Process solution ponds; 

• SX/EW plant and reagent tank farm; 

• Sulfuric acid plant and storage facilities; 

• Water treatment facilities 

• Fuel storage and dispensing facilities; and 

• Waste rock dump(s) 

Due to processing timeframes, a WPCP application should be submitted at least 180 days prior to 

the planned construction date of any component of a mining operation or the planned start of 

mining. This time frame includes the public notice and a 30-day public review and comment 

period. A WPCP is valid for 5 years, provided the operator is in compliance with the regulations. 

The Reclamation Branch of the BMRR issues a Reclamation Permit to an operator prior to 

construction of an exploration, mining, milling or other beneficiation process activity that 

proposes to disturb over five acres or remove more than 36,500 tons of material. As noted above, 

the Reclamation Permit is issued in coordination with the BLM PoO. 

Air quality permits are issued by the Bureau of Air Pollution Control (BAPC), while waterrelated 

issues (e.g., storm water discharges, sanitary septic systems, and underground injection 

control) are generally regulated by the Bureau of Water Pollution Control (BWPC). As part of 

the air permitting process, the project's potential to emit (PTE) is reviewed to determine whether 

it constitutes a major stationary source. A major stationary source is defined as either one of the 

sources identified in 40 CFR § 52.21 (including hydrofluoric, sulfuric or nitric acid plants) and 

which has a PTE of 100 tons or more per year of any regulated pollutant, or any other stationary 

source which has the PTE of 250 tons or more per year of a regulated pollutant. Based on these 

thresholds, the MacArthur Copper Project (with its sulfuric acid plant) will likely be classified as 

a major source of air pollutants that would require a Prevention of Significant Deterioration 

(PSD) and Class I air quality permit. This permit generally requires enough time to collect 

ambient air quality data and conduct detailed modeling, and will run concurrent with the 

development of the Plan of Operations and NEPA process, but would not likely be the critical 

path for the overall permitting program. 

Water appropriations, which will be important to the MacArthur Copper Project given the 

hydrologic groundwater basin in which the operations area will be located (No. 108 – Mason 

Valley) which has been “designated” with preferred uses of commercial, industrial, stock water, 

and mining, are handled through the Nevada Division of Water Resources (NDWR) and the 

State Engineer’s Office. Quaterra controls approximately 8,700 acre-feet per year (2.8 billion 

gallons per year) of appropriated water rights for mineral extraction and processing in the 

Yerington District. Some of these rights date back to the 1950’s when Anaconda operated the 

Yerington Mine. Preliminary estimates that approximately 40 percent of this water right will be 

required for the MacArthur Project. 

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20.2.3 Local Permitting 

A Special Use Permit must be acquired from Lyon County; typically a copy of the Plan of 

Operations is sufficient information for the county to review and issue this permit, although 

some additional studies may be requested, (e.g., traffic study, noise and lighting studies). 

However, these would also be addressed in the EIS. 

In addition, under county code, Title 10 – Land Use Regulations, Chapter 13 – Lyon County 

Interim Plan for Federally-Managed Public Lands, Lyon County “recognizes that the 

development of its abundant mineral resources is desirable and necessary to the state and the 

nation. Therefore, it is the policy of Lyon County to encourage mineral exploration and 

development consistent with custom and culture and to eliminate unreasonable barriers to such 

exploration and development, except for those that arise naturally from a recognition of secured 

private property rights and free market conditions.” 

20.3 ENVIRONMENTAL STUDIES 

In 2009, Quaterra expanded its Notice-level (NVN-83324) mineral exploration activities on the 

MacArthur site to include additional drilling as well as bulk sampling and up to 200 acres of 

additional surface disturbance. An exploration Plan of Operations and Reclamation Permit 

application was submitted to the BLM and NDEP, respectively, which required analysis under 

NEPA. A Plan authorization and Permit for Reclamation (Record Number NVN 

085212/Reclamation Permit No. 0294) was received in August 2009. As part of this process, a 

number of environmental baseline studies were performed to characterize the existing conditions 

within the project boundary. Much of the existing MacArthur Project site includes previously 

disturbed lands that were part of the Arimetco operations. As such, the baseline updates were 

focused on undisturbed areas. 

Vegetation, sensitive plant, weed inventories and a Class III cultural resources inventory were 

conducted in 2009. The findings were submitted to the BLM as independent baseline reports. An 

EA (DOI-BLM-NV-C020-2010-0001-EA) disclosing the potential environmental impacts 

associated with the expanded MacArthur exploration program was also published in October 

2009. A supplemental inventory was carried out in May 2012 for the areas identified for mine 

facilities in this PEA; only the sand cholla (Grusonia pulchella), a BLM special status species, 

was found within the identified project area. Mitigation of impacts to this species may include 

simply relocating the individual cacti to other locations. No federally-listed (Threatened and 

Endangered) wildlife or plant species are known to occur in the project area. 

There are no perennial surface water sources within the project area; therefore, foraging or 

incidental use for BLM sensitive bat species would be limited. Mule deer and pronghorn 

antelope distribution exist within the project area. There are no known distributions of bighorn 

sheep (Ovis Canadensis), a BLM special status species within the project area. 

To date, there have been no hydrogeological investigations or geochemical characterization 

programs for the ore, waste rock and spent leach materials, performed for the MacArthur Copper 

Project. These will be important studies to both the Plan of Operations and Water Pollution 

Control Permit, and are planned to be initiated by Quaterra as soon as practicable. Arimetco 

installed a supply-water well in 1993 at the eastern end of the property located approximately 

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4,000 ft east of the proposed mine open pits. In 2011, Quaterra rehabilitated this well and 

installed a new pump for use at the mine and in the upcoming hydrogeological investigations. 

In summary, at this time, there are no known environmental issues that would be expected to 

materially impact Quaterra’s ability to construct or operate the MacArthur Copper Project. 

20.4 WASTE AND TAILINGS DISPOSAL 

As part of both the State Water Pollution Control Permit and the BLM Plan of Operations (PoO), 

Quaterra will submit a detailed monitoring plan for monitoring to demonstrate compliance with 

the approved PoO and other federal or state environmental regulations, to provide early detection 

of potential problems, and to assist in directing potential corrective actions, should they become 

necessary. Areas of likely monitoring (particularly water monitoring) in the MacArthur Copper 

Project include: all process solutions; groundwater upgradient and downgradient of the process 

facilities (acid leach pad, solution ponds, acid plant, SX/EW); liner leak detection on the process 

ponds; and cooling tower blowdown. 

The site-wide monitoring plan will include a discussion on area water quality; monitoring 

locations, analytical profiles (NDEP Profiles I, II, or III), and sampling/reporting frequency. 

Typical monitoring programs include surface- and groundwater quality and quantity, air quality, 

revegetation, stability, noise levels, and wildlife mortality. 

The State of Nevada, through the Bureau of Mining Regulation and Reclamation (BMRR) will 

require a process fluid management plan as part of the Water Pollution Control Permit. This plan 

will describe the management of process fluids including the heap leach pad, process ponds, acid 

plant, and SX/EW plant. The plan will also provide a description of the means to evaluate the 

conditions in the fluid management system, so as to be able to quantify the available storage 

capacity for meteoric waters. 

The management of non-process (non-contact) stormwater around and between process facilities 

is a necessary part of the Nevada General Permit for Stormwater Discharges Associated with 

Industrial Activity from Metals Mining Activities (NVR300000), and is typically part of the sitewide 

Stormwater Pollution Prevention Plan (SWPPP). 

20.5 PROJECT PERMITTING REQUIREMENTS 

A detailed discussion of the project permitting requirements is provided under Section 20.2 of 

this report (above). Because of the land position of the project, both state and federal approvals 

will be required for the MacArthur Copper Project. No mine permits have thus far been acquired, 

though the appropriate permits for have been obtained under an exploration Plan of Operations. 

Bonding requirements for the operations are provided under Section 20.7 (below). 

20.6 SOCIAL OR COMMUNITY RELATED REQUIREMENTS 

Both the BLM NEPA EIS and the Lyon County SUP consider the socioeconomic impacts of a 

project prior to authorization. The MacArthur Copper Project workforce (including shorter-term 

construction contractors) will reside mainly in the town of Yerington and the surrounding 

communities in Lyon County, and possibly Storey, Douglas and Mineral counties as well. The 

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project proponent will coordinate closely with local government and businesses to ensure that the 

needs of both the community and the workforce are being met. According to the Nevada State 

Demographer, the population of Lyon County was 51,980 in 2010, up from 34,501 in 2000. This 

population growth has been slow, but steady, mainly because of an increase in agriculture and 

mining activity in the area. 

An important part of the income of predominantly rural counties in Nevada, like Lyon County, is 

produced by sales tax and the net proceeds tax on mining activity within the county. Sales tax 

revenues are collected by the county in which delivery of the goods are taken. For the MacArthur 

Copper Project, this would be Lyon County. The median household income in the county rose 

from $40,699 in 1999 to $47,518 in 2009, but is 7% below the current Nevada median income. 

In 2010, there were less than 500 persons employed in agriculture, forestry, fishing and hunting, 

and mining in Lyon County. 

20.7 MINE CLOSURE REQUIREMENTS 

Both the BLM’s 43 CFR 3809 and State of Nevada’s mine reclamation regulations require 

closure and reclamation for the MacArthur Project. In addition, any operator who conducts 

mining operations under an approved BLM PoO or State Reclamation Permit shall furnish a 

financial surety (bond) in an amount sufficient for stabilizing and reclaiming all areas disturbed 

by the operations. 

In general, buildings and facilities not identified for a post-mining use will be removed from the 

site during the salvage and site demolition phase. Above-ground concrete will be demolished and 

removed from site or buried on site. Below-ground concrete will remain and be covered. 

Residual solution remaining in heap leach pad and process circuit will be recirculated until the 

rate of flow from these facilities can be passively managed through evaporation from the ponds 

or a combination of evaporation and infiltration. The heap leach pad and mine waste dumps will 

be re-contoured to a 3:1 slope, covered with available growth media, and revegetated. 

Reclamation and closure activities will be conducted concurrently, to the extent practical, to 

reduce the overall reclamation and closure costs, minimize environmental liabilities, and limit 

bond exposure. 

The revegetation release criteria for reclaimed areas are presented in the “Guidelines for 

Successful Revegetation for the Nevada Division of Environmental Protection, the Bureau of 

Land Management, and the U.S.D.A. Forest Service.” The revegetation goal is to achieve the 

permitted plant cover as soon as possible. 

Conceptual reclamation and closure methods were used to evaluate the various components of 

the project to estimate reclamation costs. Quantities were estimated based on the physical layout, 

geometry and dimensions of the proposed project components of the site plan and facilities 

layout. These included current conceptual designs for the main project components including the 

open pit, infrastructure, waste rock facilities, acid leach pad, and process ponds. Equipment and 

labor costs were also estimated based on current industry rates. A 20-percent contingency was 

applied to this estimate. 

Because the closure activities for the PEA are based on preliminary designs and conceptual 

approaches, the overall closure cost estimate is considered to be conservative. The closure cost 

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associated with the MacArthur Copper Project is currently estimated to be $92 million ($82 

million after salvage). This total is an undiscounted internal cost to reclaim and close the 

facilities associated with the mining and processing project. 

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21 CAPITAL AND OPERATING COSTS 

21.1 CAPITAL COST 

21.1.1 Mine Capital Cost 

The mine capital cost estimate was provided by Independent Mining Consultants (IMC) and is 

estimated to be $48 million for the initial capital and $83.6 million in sustaining capital. The 

initial and sustaining capital consists of the initial fleet of mining and support equipment, with 

additions to the fleet as necessary. Replacement of some of the equipment fleet is also included 

in the sustaining capital. 

21.1.2 SX/EW Capital Cost 

The total installed capital cost for the SX/EW and ancillary facilities is estimated to be $114.3 

million and is summarized by process area in Table 21-1 below. 

Table 21-1: SX/EW Capital Cost 

Direct Field Cost $ 000 

000 Plant General $1,039 

300 Heap Leach Pad $17,368 

350 Solutin Ponds $7,767 

400 Solvent Extraction $10,918 

500 Tank Farm $10,359 

600 Electrowinning $16,263 

650 Water Systems $1,309 

700 Main Substation $2,499 

750 Transmission Line $48 

800 Reagents $1,160 

900 Ancillary Facilities $3,963 

$72,693 

Indirect Cost 

Mobilization $727 

Lyon County Sales Tax $3,740 

Freight $4,743 

EPCM $12,932 

Vendor Supervision & Commissioning $423 

Contingency (20%) $19,052 

Total Direct and Indirect Capital Cost $114,310 

The initial capital cost is based on recent M3 Engineering & Technology in-house data and 

previous estimates for SX/EW facilities of similar size. The construction labor was adjusted to 

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Davis-Bacon March 2012 prevailing shop wages in Lyon County, Nevada and construction 

materials and equipment were factored as required based on PLS flow rates or total copper 

production to arrive at a total direct capital cost. Indirect capital costs were developed from the 

direct field cost based on in-house factors. Indirect field mobilization is 1.0% of the direct field 

cost; Lyon County sales tax is 7.1% of direct field cost less labor; freight is 10% of total 

equipment and materials; engineering, procurement and construction management (EPCM) is 

16.7% of the direct field cost plus the indirect costs listed above; commissioning, commissioning 

spares, and vendor pre-commissioning and supervision is 3.1% of the plant equipment cost; and 

a contingency of 20% was applied. The accuracy range of the estimate is -20% to +25%, suitable 

to support a Preliminary Economic Assessment. 

Sustaining capital for the SX/EW and heap leach pad is summarized in Table 21-2 below and 

includes expansions of the heap leach pad and replacement of the mobile process equipment. 

Normal maintenance and repair of process equipment is included as part of the operating cost. 

Table 21-2: SX/EW Sustaining Capital 

SX/EW 

Mobile 

Equipment 

Heap Leach 

Phases Total 

Year $000 $000 $000 

3 $5,383 $5,383 

5 $50 $31,439 $31,489 

6 $50 $50 

7 $51 $8,689 $8,740 

8 $151 $151 

9 $135 $135 

10 $316 $5,346 $5,662 

11 $250 $250 

12 $76 $3,344 $3,420 

13 $76 $3,749 $3,825 

14 $50 $4,812 $4,862 

$1,205 $62,762 $63,967 

The process areas making up the initial capital cost estimate for the MacArthur SX/EW facility 

are defined below. 

Site General (Area 000) 

The Site General Area consists of systems or facilities that cross multiple areas of the plant. This 

area consists of the overall site grading, internal access roads, perimeter fencing, and 

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instrumentation software, licenses and programming. Since the project is adjacent to existing 

roads and infrastructure, there are no costs required for main access roads to the property. 

Heap Leach Pad (Area 300) 

This area consists of the site grading for the initial phase heap leach pad, a GCL liner, and a 60 

mil LLDPE liner with an anchor trench around the perimeter to anchor the lining. Also included 

is a perforated HDPE piping network to collect the leach solution and the crushing and 

placement of over-liner material to protect the collection piping system. Raffinate distribution 

piping on top of the heap leach pad is also included. The initial heap leach pad covers 

approximately 123 acres. The development of the remaining area of the heap leach pad (~264 

acres) is included in the sustaining capital. 

Solution Ponds (Area 350) 

This area consists of the pregnant leach solution (PLS) pond, raffinate pond and a storm water 

collection pond. The PLS and raffinate ponds are double lined with HDPE liners and a leak 

detection system between liners. The storm water event pond is lined with a single HDPE liner. 

After a storm event, any solution flowing into the event pond will be pumped to the raffinate 

pond. All solution ponds are fenced. The solution piping between the heap leach pad and the 

solvent extraction facility is also included in this area. 

Solvent Extraction (Area 400) 

This area consists of four extraction settlers and one stripping settler, including primary and 

secondary mix tanks and agitators. An SX feed tank is included to provide a consistent gravity 

feed to the extraction settlers. PLS is pumped from the PLS pond to the SX feed tank and then by 

gravity to the extraction settlers. The copper depleted raffinate will flow by gravity from the 

extraction settlers to the raffinate pond. A foam fire suppression system is provided in this area 

for fire suppression. 

Tank Farm (Area 500) 

This area contains the circulation tanks, pumps, heat exchangers, and filters that support the 

solvent extraction and electrowinning facilities. Included in this area are the loaded organic 

tanks, electrolyte circulating tanks, electrolyte filters, electrolyte heat exchangers, and a diluent 

storage tank. A crud holding tank and crud recovery equipment is also included. 

Electrowinning Facility (Area 600) 

This area includes the electrowinning tank house with 54 electrolytic cells, cathode and anode 

electrodes, a transformer / rectifier, a semi-automatic cathode stripping machine and a boiler to 

provide hot water for cathode washing and for maintaining heat in the circulating electrolyte 

system. Also included is a tank house ventilation system and scrubber for tank house mist 

control. 

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Fresh Water System (Area 650) 

This area consists of fresh water wells located on site; a combined fresh water and fire water 

storage tank; a potable water treatment, storage, and distribution system; and the fire water 

pumps and distribution system. Fresh water will be pumped from onsite wells to the fresh water 

storage tank and distributed to all areas of the plant. 

Main Electrical Substation (Area 700) 

This area consists of the main electrical substation, switchgear and transformers to transform 

electrical power from the 69 kV main line to intermediate voltages for distribution throughout 

the plant site. The cost for motor control centers in various areas of the plant, along with the low 

voltage distribution to the end users, is included in the process area cost. 

Power Transmission Line (Area 750) 

This area includes a dead end structure and tap at an existing 69 kV main power line and a short 

overhead line to the plant main electrical substation. The main power line feeds from the Fort 

Churchill power plant, owned by Nevada Energy, to the town of Yerington and runs adjacent to 

the SX/EW location. An allowance is provided for minimum upgrades to the main power line in 

the area of the connection. 

Reagents (Area 800) 

The reagents area consists of receiving, storage and distribution of reagents used in the SX/EW 

process. Reagents include the SX extractant, diluent (kerosene) for the organic, cobalt sulfate 

and guar in the tank house, and mist suppressor (FC-1100) to suppress acid mist in the tank 

house. Sulfuric acid for the leaching process will be supplied by an onsite sulfuric acid plant 

discussed in Section 2.1.3. 

Ancillary Facilities (Area 900) 

Ancillary facilities provided for the project include a change house, a modular guard house and 

truck scale at the plant entrance, and fuel storage and dispensing facilities for diesel fuel and 

gasoline. Allowances have also been provided to upgrade existing buildings at the Yerington 

property to be used for an administration building, warehouse, analytical laboratory, maintenance 

building, and mine truck shop. Powder and detonator magazines are assumed to be provided by 

the explosives supplier in exchange for a long term supply contract. 

21.1.3 Sulfuric Acid Plant Capital Cost 

The base case for the MacArthur Project considers an onsite sulfuric acid plant sized for 640 tons 

per day of sulfuric acid, in accordance with the expected acid consumption from the leaching 

operation. The total installed capital cost for the sulfuric acid plant is estimated to be $65.4 

million and is summarized in Table 21-3 below. 

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Table 21-3: Sulfuric Acid Plant Capital Cost 

$000 

Area Direct Cost 

810 Sulfur Unloading $5,563 

820 Acid Plant $28,065 

830 Acid Storage $1,638 

840 Power Plant $6,155 

850 Water Treatment $1,344 

860 Cooling Towers $921 

Total Direct Cost $43,687 

Indirect Cost 



EPCM $7,732 

Vendor Supervision & Commissioning $500 


Total Direct and Indirect Cost $65,439 

The initial capital cost is based on recent M3 Engineering & Technology in-house data on 

previous sulfuric acid plant facilities. Construction labor was adjusted for January 2012 Davis- 

Bacon prevailing shop wages in Lyon County, Nevada. The direct field costs were factored 

based on the acid plant capacities. As with the other facilities, indirect capital costs were 

developed from the direct field cost based on in-house factors. Indirect field mobilization is 

1.0% of the direct field cost; Lyon County sales tax is 7.1% off direct field cost less labor; freight 

is 10% of total equipment and materials; engineering, procurement and construction management 

(EPCM) is 16.7% of the direct field cost plus the indirect costs listed above; commissioning, 

commissioning spares, and vendor pre-commissioning and supervision is 3.1% of the plant 

equipment cost; a contingency of 20% was applied. The accuracy range of the estimate is -20% 

to +25%, suitable to support a Preliminary Economic Assessment. 

The capital cost estimate for the sulfuric acid plant and associated facilities is an additional cost 

to the estimate for the SX/EW facilities. Common facilities already included in the SX/EW 

estimate are not included in the sulfuric acid plant estimate. 

Sustaining capital for the sulfuric acid plant and associated facilities was not estimated; however, 

an accrual is included in the operating and maintenance cost for major repairs required at 

intervals of 1.5 to 2 years. 

The process areas that make up the direct capital cost for the sulfuric acid plant are defined 

below: 

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Sulfur Handling and Unloading (Area 810) 

This area consists of facilities to receive molten sulfur by rail tank cars, unloading to receiving 

pits, and pumping to heated storage. Also included is a direct fired boiler used to generate steam 

for heating the rail cars for unloading and maintaining heat in the sulfur tanks and pipelines. 

These facilities are to be located at an existing rail siding near Wabuska, Nevada, approximately 

eight miles from the project site. The molten sulfur will be transferred by truck from the 

unloading location to the molten sulfur storage tanks at the acid plant. Trucks are provided in the 

capital cost estimate for this transfer. 

Sulfuric Acid Plant (Area 820) 

The sulfuric acid plant is a double-absorption double-contact plant and consists of the sulfur 

burning furnace, a waste heat boiler to cool the combustion gases and generate steam, a main gas 

blower to provide dry combustion air to the sulfur furnace and deliver the combustion gas 

through the sulfuric acid plant, and a converter with associated heat exchangers to convert SO2 

to SO3 in the combustion gas. Also included are a drying tower, intermediate absorption tower 

and final absorption tower with associated acid pump tanks and acid coolers. Final waste gas 

from the final absorption tower is vented to atmosphere through a final tail gas stack. 

Sulfuric Acid Storage (Area 830) 

This area consists of sulfuric acid storage tanks for the product acid from the acid plant. Sulfuric 

acid will be pumped or gravity fed to the raffinate pond and SX plant for use in the leach 

operation. 

Power Plant (Area 840) 

The power plant includes the steam turbine generator, main condenser, dump condenser, and 

steam separator. Cooling water for the steam condensers will be provided by the cooling towers 

in Area 860. The power generated by the steam turbine will be connected with the main buss at 

the SX/EW main substation for distribution to all areas of the plant. The breakers and 

synchronizing equipment to connect to the main buss is included in this area. 

Water Treatment (Area 850) 

This area includes the water treatment facilities to produce boiler quality water. The treatment 

facility includes fresh water filters, water softeners, Reverse osmosis (RO) filters, oxygen 

scavengers, boiler feed water pumps and tanks. Chemical water treatment for the cooling towers 

is included with the cooling towers in Area 860. 

Cooling Towers (Area 860) 

This area consists of the cooling towers, fans, circulating water pumps, and a chemical water 

treatment system. The cooling tower serves the sulfuric acid plant and the power plant. The 

cooling towers are located adjacent to the sulfuric acid plant. 

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21.1.4 Exclusions 

Owner’s costs have been excluded from the capital cost estimate; however, an allowance of $5 

million has been included in the financial analysis for typical Owner’s costs such as first fills of 

reagents and lubricants, office equipment, and Owner’s pre-production staffing. 

The Owner’s costs noted below are not included in the $5.0 million allowance. 

a) Environmental permits, 

b) Performance bond, 

c) Builder’s risk insurance, 

d) Land Acquisition, 

e) Water rights acquisition, 

f) Sunk costs prior to the estimate, and 

g) Escalation, 

21.2 RECLAMATION COST ESTIMATE 

A capital cost estimate was prepared for reclamation of the SX/EW and acid plant site, as well as 

the leach pad and mine waste dumps. The reclamation cost includes dismantling all buildings 

and equipment and removing from the site. Above ground concrete will be demolished and 

removed from site or buried on site. Below ground concrete will remain and be covered. 

Solution ponds will be drained and the top lining removed to inspect the bottom lining for leaks. 

If there is evidence of leaks; the bottom lining will be removed, the soil at the leak tested for 

contamination, and any required remediation performed before the pond can be covered. If no 

evidence of leaks is found, the top lining can be folded over in place and the pond covered. The 

ponds will be filled by the push down of the heap leach pad as part of the reclamation cost for the 

leach pad. A mound is provided over the pond area to prevent storm water from collecting over 

the pond and migrating into the pond. The plant site will be graded to approximate original 

contours. Roads will be left in place; however, any asphalt will be removed. The leach pad and 

mine waste dumps will be re-contoured to a 3.5: 1 slope, with setbacks, covered with reclaimed 

soil, and hydro-seeded for plant growth. The area of the SX/EW and sulfuric acid plant will also 

be hydro-seeded for plant growth. 

The indirect cost for the reclamation estimate includes field mobilization at 1% of the direct field 

cost, Lyon County sales tax at 7.1% of direct field cost less labor, and a contingency of 20%. It is 

assumed the management of the reclamation effort will be by the Owner’s team already on site; 

therefore, no allowance is provided for EPCM. 

The total cost for reclamation of the site is estimated to be $92.2 million and is summarized by 

process area in Table 21-4 below. The cost to re-contour the mine waste dumps is included in 

Area 300 with the cost to re-contour the heap leach pad. 

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Table 21-4: Reclamation Cost Estimate 

Direct Field Cost $000 

000 Plant General $53 

300 Heap leach $58,067 

350 Solution Ponds $868 

400 Solvent Extraction $1,772 

500 Tank Farm $1,655 

600 Electrowinning $2,339 

650 Water Systems $137 

700 Main Substation $237 

750 Overhead Transmission line $8 

800 Reagents & Acid Plant $8,840 

900 Ancillaries $115 

Total Direct Field Cost $74,091 

Indirect Costs 




Total Indirect Cost $92,159 

An allowance of $9.2 million was provided for equipment and materials salvage to offset the 

reclamation cost. Total reclamation cost with the salvage deduct is $82.96 million, which occurs 

in years 19 through 22. 

21.3 OPERATING COST 

The overall annual average operating cost for the MacArthur Copper operation is $1.89 per 

pound of copper and is summarized in Table 21-5 below. The costs include the mine operations, 

SX/EW facility, the sulfuric acid plant, general administrative expenses and the cost of 

transportation to ship the product cathode to market. The sulfuric acid operating cost represents 

the cost of acid in the table below. 

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21.3.1 Mine Operating Cost 

Table 21-5: MacArthur SX/EW and Mine Operating Cost 

$ / lb. Cu 

Mine $0.99 

SX/EW $0.38 

Cost of Acid $0.35 

General & Administrative $0.12 

Transportation $0.05 

Sub-Total $1.89 

The mine operating costs were provided by Independent Mining Consultants (IMC) based on a 

selected fleet of mine and support equipment for the 18 year life of the MacArthur mine. The 

average life of mine operating cost is $1.44 per ton of material mined or $2.74 per ton of ore 

mined. The mine operating cost, as a cost per pound of copper recovered, is $0.99 per pound of 

copper produced. These costs include drilling, blasting, loading, hauling, and road and dump 

maintenance. 

21.3.2 SX/EW Operating Cost 

The operating cost estimate for the MacArthur SX/EW facilities is estimated to be $0.38 per 

pound of copper produced and include labor, reagents, electrical power, maintenance parts and 

services and operating supplies and services. The costs are summarized in Table 21-6 below. 

Table 21-6: SX/EW Operating Cost 

$ / lb. Cu 

Labor $0.08 

Reagents $0.10 

Electrical Power $0.14 

Maintenance Parts & Services $0.03 

Operating Supplies & Services $0.03 

Sub-Total $0.38 

The average annual labor cost for the SX/EW area is $3.5 million based on a staffing plan of 43 

operating and maintenance personnel. The average wage rate for the SX/EW staff is $58,600 per 

year plus 40% for fringe benefits. The labor staffing consists of four supervisory personnel, 

twenty-two operating personnel and seventeen maintenance personnel. 

The average annual cost of reagents for this area is $4.15 million and includes extractant, diluent, 

cobalt sulfate, and guar. The cost of acid is noted separately and is based on the operating costs 

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of the onsite sulfuric acid plant discussed in Section 21.3.3 below. The annual consumption of 

SX/EW reagents and cost is shown in Table 21-7 below. 

Table 21-7: Reagent Cost 

Consumption $ / lb. Cu 

Extractant 106 gallons / day $0.032 

Diluent 1,415 gallons / day $0.053 

Cobalt Sulfate 163 pounds /day $0.007 

Guar 29.5 pounds / day $0.005 

Total $0.097 

The annual power cost for the SX/EW facility is $5.7 million, or $0.14 per pound of copper 

recovered, and is based on a power consumption of approximately 2.1 kWh per lb. of cathode 

copper and a cost of power of $0.065 /kWh. 

The annual cost for maintenance parts and services is approximately $1.4 million and is based on 

7% of the SX/EW equipment cost. The annual cost of operating supplies and services is $1.2 

million and is based on $0.03 / lb. of cathode copper produced. 

21.3.3 Sulfuric Acid Plant Operating Cost 

The annual operating cost for the sulfuric acid plant, power plant and associated facilities is 

$14.4 million or $62.04 per ton of acid and $0.35 per pound of copper produced. The estimated 

cost for sulfuric acid delivered to site by rail from the west coast is estimated to be $140 per ton 

of acid compared to the cost to manufacture on site at $62 per ton of acid. The acid plant 

operating costs are summarized in Table 21-8 below. 

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Table 21-8: Sulfuric Acid Plant Operating Cost 

Annual Cost Cost / Cost / lb. 

Cost ton Acid Copper 

$000 

Labor $1,401 $6.03 $0.03 

Reagents (Sulfur) $9,512 $40.95 $0.23 

Fuels (Propane) $634 $2.73 $0.02 

Power (Credit) ($2,624) ($11.30) ($0.06) 

Maintenance $3,846 $16.56 $0.09 

Operating Supplies $1,643 $7.07 $0.04 

$14,412 $62.04 $0.35 

The labor cost is based on a staffing plan of 10 operators and 7 maintenance personnel. The 

operating crew consists of a general foreman and technician on day shift, 5 days per week and a 

control room operator and field operator each shift seven days per week. The average annual 

wage rate for acid plant personnel is $58,800 plus 40% fringe benefits. The wage rate is slightly 

higher in the acid plant than the SX/EW facility due to the higher mix of higher pay positions in 

the acid plant. 

Reagents needed in the sulfuric acid plant includes elemental sulfur (molten) for acid production 

and water treatment chemicals for the cooling tower and boiler feed water systems. One ton of 

sulfur will produce a little over 3 tons of sulfuric acid. Based on an annual requirement of 

232,300 tons of sulfuric acid, approximately 77,400 tons of elemental sulfur will be required. 

The cost of sulfur used in the estimate is $125 per ton delivered molten to site and is based on the 

average cost for U. S. West Coast sulfur over the last five years of available published 

information with freight allowed to the project site. An allowance of $30,000 per year was used 

for the water treatment chemicals. 

Propane is assumed as the fuel to fire the steam boiler at the sulfur unloading area and is based 

on a boiler sized for 5 million BTU/hr and a heat value for Propane of 92,500 BTU/gallon. It is 

assumed that the boiler would operate 16 hours per day. The cost of Propane was set at $2.00 per 

gallon, the average of current wholesale and residential cost. 

The power requirement to produce sulfuric acid was estimated to be 2,300 kW or $1.3 million 

annually at the cost of power of $0.065 per kWh. The turbine generator is expected to produce 

approximately 6,900 kW of power at a value of $3.9 million annually at the same cost of power. 

The excess power can be used to displace purchased power in the SX/EW facility or sold back to 

the power company. The net power credit is $2.6 million annually. The power consumption and 

power produced were factored from existing in-hours data on similar sulfur burning acid plants. 

Annual maintenance cost for the sulfuric acid plant was estimated at 4% of the installed cost of 

the acid plant, or $2.6 million. The annual maintenance for the power plant was estimated to be 

$0.02 / kWh or $1.2 million. The total annual maintenance cost for the acid plant and power 

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plant is $3.8 million. The maintenance cost includes an accrual for major repairs that will occur 

at intervals of 1.5 to 2 years. 

Operating supplies and services was estimated at 2.5% if the total installed cost of the acid plant 

and power plant or $1.6 million annually. 

21.3.4 General and Administrative Costs 

The total annual general and administrative cost for the facility is $5.0 million, or $0.12 per 

pound of cathode copper produced. The G&A labor is the largest component at $2.2 million per 

year, based on a staffing of 23 employees. Allowances were made for non-labor components for 

G&A expenses, which includes office supplies, fuels, communications, small vehicle 

maintenance, claims assessments, legal and auditing, insurance, travel, meals and expenses, 

community relations, recruiting and relocation expenses, and janitorial services. The breakdown 

of G&A cost and labor detail is shown in Table 21-9 General & Administrative Cost Summary 

and Table 21-10 General & Administrative Labor Cost Summary. 

Table 21-9: General & Administrative Cost Summary 

Cathode Produced (lbs.) 41,500,000 

Total 

Cost Item Annual Cost - $ $/lb. Cathode 

Labor & Fringes $2,186,800 $0.053 

Accounting (excluding labor) $25,000 $0.001 

Safety & Environmental (excluding labor) $25,000 $0.001 

Human Resources (excluding labor) $25,000 $0.001 

Security (excluding labor) $25,000 $0.001 

Office Operating Supplies and Postage $40,000 $0.001 

Fuel/ Propane $25,000 $0.001 

Communications $70,000 $0.002 

Small Vehicles $125,000 $0.003 

Claims Assessment $10,000 $0.000 

Legal & Audit $300,000 $0.007 

Consultants $250,000 $0.006 

Janitorial Services $50,000 $0.001 

Insurances $1,000,000 $0.024 

Taxes - property $500,000 $0.012 

Subs, Dues, PR, and Donations $60,000 $0.001 

Travel, Lodging, and Meals $150,000 $0.004 

Recruiting/Relocation $125,000 $0.003 

Total General & Administrative Cost $4,991,800 $0.120 

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Table 21-10: General & Administrative Labor Cost Summary 

Department Number Total Annual Total Annual 

General & Administrative 

and Of Direct Benefits Total Annual 

Position Personnel Salary 40% Salary 

General Manager 1 $200,000 $80,000 $280,000 

Administrative Assistant 1 $34,000 $13,600 $47,600 

Controller 1 $120,000 $48,000 $168,000 

Accountant 1 $60,000 $24,000 $84,000 

Accounts Payable 1 $34,000 $13,600 $47,600 

Purchasing Manager 1 $120,000 $48,000 $168,000 

Purchasing Agent 1 $55,000 $22,000 $77,000 

Warehouseman 2 $40,000 $16,000 $112,000 

IT Technician 2 $60,000 $24,000 $168,000 

HR Manager 1 $120,000 $48,000 $168,000 

HR Specialist 1 $40,000 $16,000 $56,000 

HR Administrative Assistant 1 $34,000 $13,600 $47,600 

Safety Manager 1 $120,000 $48,000 $168,000 

Environmental Manager 1 $120,000 $48,000 $168,000 

Safety Specialist 2 $55,000 $22,000 $154,000 

Hydrological Engineer 0 $60,000 $24,000 $0 

Environmental Technician 1 $55,000 $22,000 $77,000 

Security Guard 4 $35,000 $14,000 $196,000 

Total General and Administrative 

Labor Cost 

23 $2,186,800 

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$0



22 ECONOMIC ANALYSIS 

22.1 INTRODUCTION 

The financial evaluation presents the determination of the Net Present Value (NPV), payback 

period (time in years to recapture the initial capital investment), and the Internal Rate of Return 

(IRR) for the project. Annual cash flow projections were estimated over the life of the mine 

based on the estimates of capital expenditures and production cost and sales revenue. The sales 

revenue is based on the production of copper cathode. The estimates of capital expenditures and 

site production costs have been developed specifically for this project and have been presented in 

earlier sections of this report. 

22.2 MINE PRODUCTION STATISTICS 

Mine production is reported as ore and overburden from the mining operation. The annual 

production figures were obtained from the mine plan as reported earlier in this report. The life of 

mine ore quantities and ore grades are presented in Table 22-1 below. 

Table 22-1: Life of Mine Ore, Waste Quantities, and Ore Grade 

Oxide Ore – Main Pit 

Oxide Ore – Other Areas 

Mixed Ore 

Tons 

(kt) 

132,756 

52,537 

85,588 


% 

0.20 

0.19 

0.24 

Total Ore 270,881 0.21 

Waste 244,948 

22.3 HEAP LEACH PAD AND SX/EW PRODUCTION STATISTICS 

The ore will be processed using heap leaching and SX/EW plant recovery technology to produce 

copper cathode. Below are the recoveries assigned to each of the ore types: 

Oxide Ore – Main Pit 70.0% 

Oxide Ore – Other Areas 65.0% 

Mixed Ore 60.0% 

The estimated life of mine metal production is estimated to be 747.7 million pounds. 

22.3.1 Cathode Shipping 

The cost for cathode shipping of $0.05 per pound of copper is included in cash operating costs. 

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22.4 CAPITAL EXPENDITURE 

22.4.1 Initial Capital 

The financial indicators have been determined with 100% equity financing of the initial capital. 

Any acquisition cost or expenditures prior to start of mine pre-development have been treated as 

“sunk” cost and have not been included in the analysis. 

The total initial capital carried in the financial model for new construction is expended over a 3 

year period. The initial capital includes Owner’s costs and contingency. The cash flow will be 

expended in the years before production and a small amount carried over into the first production 

year. 

The initial capital is presented in Table 22-2 below. 

Table 22-2: Initial Capital 

$ in millions 

Mining $48.0 

SXEW Plant $114.3 

Sulfuric Acid Plant $65.4 

Owner's Cost $5.0 

Total 

22.4.2 Sustaining Capital 

$232.7 

A schedule of capital cost expenditures during the production period was estimated and included 

in the financial analysis under the category of sustaining capital. The total life of mine sustaining 

capital is estimated to be $147.6 million. This capital will be expended during a 14 year period 

during the 18 year mine life. 

22.4.3 Working Capital 

A 15 day delay of receipt of revenue from sales is used for accounts receivables. A delay of 

payment for accounts payable of 30 days is also incorporated into the financial model. In 

addition, working capital allowance of $1.1 million for plant consumable inventory is estimated 

in year -1 and year 1. All the working capital is recaptured at the end of the mine life and the 

final value of these accounts is $0. 

22.4.4 Salvage Value 

An allowance based on 10% of the total capital equipment cost, including mine equipment for 

salvage value at the end of the mine life has been included and is estimated at $9.2 million. 

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22.5 REVENUE 

Annual revenue is determined by applying estimated copper price to the annual payable metal 

estimated for each operating year. Sales prices have been applied to all life of mine production 

without escalation or hedging. The revenue is the gross value of payable metals sold before 

treatment charges and transportation charges. The copper sales price used in the evaluation is 

based on the three year historical price as of May 1, 2012, which is consistent with Securities and 

Exchange Commission (SEC) guidelines. 

Copper $3.48/pound 

22.6 OPERATING COST 

The average Cash Operating Cost over the life of the mine is estimated to be $1.89 per pound of 

copper, excluding the cost of the capitalized pre-stripping. Cash Operating Cost includes mine 

operations, process plant operations, general administrative cost, and shipping charges. Table 

22-3 below shows the estimated operating cost by area per pound of copper. 

22.7 TOTAL CASH COST 

Table 22-3: Life of Mine Operating Cost 

Operating Cost $/lb 

Mine $0.99 

SXEW Plant $0.38 

Sulfuric Acid Cost $0.35 

General Administration $0.12 

Transportation $0.05 

Total Operating Cost $1.89 

The average Total Cash Cost over the life of the mine is estimated to be $2.04 per pound of 

copper. Total Cash Cost is the Operating Cost plus royalty, salvage value, reclamation and 

closure costs. 

22.7.1 Royalty 

The royalty charges for the life of the mine are estimated at $31.3 million. There are two 

royalties and they are based on a % of net smelter return. The net smelter return is calculated as 

gross revenues, less SX/EW cost (excluding sulfuric acid cost) and transportation cost. The 

royalties are defined as follows: 

• Arimetco royalty which is based on 2% of the net smelter return and has a cap of $7.5 

million. 

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• A 3 rd Party royalty which is based on a payment of $1.0 million at the start of production plus 

1% of the net smelter return for life of mine. 

22.7.2 Reclamation and Closure 

An estimate for reclamation and closure was included in the cash flow of $92.2 million. 

22.8 DEPRECIATION AND DEPLETION 

Depreciation 

Depreciation is calculated using the MACRS method starting with first year of production. The 

initial capital and sustaining capital used a 7 year life. The last year of production is the catch-up 

year if the assets are not fully depreciated by that time. 

Depletion 

The percentage depletion method was used in the evaluation. It is determined as a percentage of 

gross income from the property, not to exceed 50% of taxable income before the depletion 

deduction. The gross income from the property is defined as metal revenues minus downstream 

costs from the mining property (smelting, refining and transportation). Taxable income is defined 

as gross income minus operating expenses, overhead expenses, and depreciation and state taxes. 

The estimated depletion deduction for income tax use is $346.2 million for the life of the mine. 

22.9 TAXATION 

22.9.1 Income Tax and Mineral Tax 

Taxable income for income tax purposes is defined as metal revenues minus operating expenses, 

royalty, property and severance taxes, reclamation and closure expense, depreciation and 

depletion. The Federal income tax rate is 35% in accordance with IRS Publication 542. 

Federal income taxes were calculated on the taxable income described above using the federal 

tax rate. Nevada does not have a state corporate income tax. 

Federal income taxes paid are estimated to be $151.7 million. 

In addition to the Federal income tax, a Nevada mineral tax has also been estimated at $27.4 

million. The Nevada Net Proceeds of Minerals Tax is an ad valorem property tax assessed on 

minerals mined or produced in Nevada when they are sold or removed from the state. 

22.10 PROJECT FINANCING 

The project was evaluated on an unleveraged and un-inflated basis. 

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22.11 NET INCOME AFTER TAX 

The operating margin before tax is $840.87 million, including depreciation, and the net income 

after tax amounts to $514.2 million. 

22.12 NPV AND IRR 

The economic analysis indicates that the project has an NPV of $201.6 million at a discount rate 

of 8%, an Internal Rate of Return (IRR) of 24.2% with a payback period of 3.1 years after taxes. 

22.13 SENSITIVITIES 

Table 22-4: Economic Indicators 

Before Taxes After Taxes 

$000 

$000 

NPV @ 8% 284,138 201,576 

IRR % 29.3% 24.2% 

Payback, years 2.7 3.1 

Table 22-5 compares the base case project after tax financial indicators with the financial 

indicators when different variables are applied. By comparing the results it can be seen that the 

copper price has the most impact on the project followed by the operating cost and then by the 

initial capital cost. This data is represented in graph form in Figure 22-1. The discounted cash 

flow model for the project is shown in Table 22-6 at the end of this section. 

Table 22-5: Sensitivity Analysis 

NPV @ 8% IRR Payback 

Base Case $201,576 24.2% 3.1 

Copper Price +20% $377,172 35.2% 2.3 

+10% $290,768 29.9% 2.7 

-10% $107,566 17.4% 3.7 

-20% $9,797 9.0% 8.4 

Capital Cost +20% $167,445 19.4% 3.6 

+10% $184,561 21.6% 3.4 

-10% $218,571 27.3% 2.8 

-20% $234,567 31.0% 2.5 

Operating Cost +20% $107,289 17.8% 3.5 

+10% $156,080 21.3% 3.3 

-10% $245,478 26.8% 2.9 

-20% $286,955 29.1% 2.8 

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Figure 22-1: MacArthur Project NPV Sensitivities 

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Table 22-6: Discounted Cash Flow Model 

Base Case with Acid Plant 

Mine Operations 

Total -2 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 

Oxide Ore - Main Pit (kt) 132,756 

- 

399 15,000 15,000 14,204 14,228 9,306 11,794 

8,181 4,369 4,405 1,190 

- 

1,202 3,440 3,670 4,400 4,639 8,809 8,489 

31 

- 

- 

- 

- 

Copper Grade % 0.198% 0.000% 0.241% 0.210% 0.239% 0.210% 0.210% 0.199% 0.190% 0.180% 0.180% 0.190% 0.190% 0.000% 0.163% 0.210% 0.180% 0.160% 0.168% 0.179% 0.170% 0.180% 0.000% 0.000% 0.000% 0.000% 

Contained Copper (klbs) 525,765 

- 

1,920 62,851 71,590 59,695 59,721 37,013 44,751 29,452 15,728 16,739 4,522 

- 

3,926 14,482 13,198 14,071 15,628 31,505 28,863 

112 

- 

- 

- 

- 

Mixed Ore (kt) 85,588 

- 

- 

- 

- 

6 

234 4,654 

57 

1,096 4,271 5,824 9,473 12,617 10,957 3,698 6,206 6,507 7,253 5,773 6,511 

451 

- 

- 

- 

- 



- 

- 

- 

- 

29 1,030 21,408 

217 

4,384 20,501 26,790 45,470 60,562 59,168 23,667 33,512 35,138 31,913 26,556 26,044 1,624 

- 

- 

- 

- 

Oxide Ore - Other Areas (kt) 52,537 

- 

- 

- 

- 

790 

538 1,040 3,149 

5,723 6,360 4,771 4,337 2,383 2,841 7,862 5,124 4,093 3,108 

418 

- 

- 

- 

- 

- 

- 



- 

- 

- 

- 

2,721 2,034 3,936 13,400 22,904 24,166 17,126 13,888 8,579 11,957 32,888 18,361 13,904 11,189 1,170 

- 

- 

- 

- 

- 

- 

Waste (kt) 244,948 

Total Material Mined (kt) 515,829 

- 

- 

101 

500 

4,042 

19,042 

2,174 

17,174 

5,000 

20,000 

5,000 

20,000 

5,000 

20,000 

15,000 

30,000 

20,000 

35,000 

20,000 

35,000 

SXEW Operations 

Oxide Ore - Main Pit (kt) 132,756 

- 

- 

15,399 15,000 14,204 14,228 9,306 11,794 

8,181 4,369 4,405 1,190 

- 

1,202 3,440 3,670 4,400 4,639 8,809 8,489 

31 

- 

- 

- 

- 

Copper Grade 0.198% 0.000% 0.000% 0.210% 0.239% 0.210% 0.210% 0.199% 0.190% 0.180% 0.180% 0.190% 0.190% 0.000% 0.163% 0.210% 0.180% 0.160% 0.168% 0.179% 0.170% 0.180% 0.000% 0.000% 0.000% 0.000% 


- 

- 

64,771 71,590 59,695 59,721 37,013 44,751 29,452 15,728 16,739 4,522 

- 

3,926 14,482 13,198 14,071 15,628 31,505 28,863 

112 

- 

- 

- 

- 

Recovery 70.0% 0.0% 0.0% 70.0% 70.0% 70.0% 70.0% 70.0% 70.0% 70.0% 70.0% 70.0% 70.0% 70.0% 70.0% 70.0% 70.0% 70.0% 70.0% 70.0% 70.0% 70.0% 0.0% 0.0% 0.0% 0.0% 

Recovered Copper (klbs) 368,036 

- 

- 

45,340 50,113 41,786 41,805 25,909 31,326 20,616 11,010 11,717 3,165 

- 

2,748 10,137 9,239 9,850 10,939 22,054 20,204 

78 

- 

- 

- 

- 

Mixed Ore (kt) 85,588 

- 

- 

- 

- 

6 

234 4,654 

57 

1,096 4,271 5,824 9,473 12,617 10,957 3,698 6,206 6,507 7,253 5,773 6,511 

451 

- 

- 

- 

- 



- 

- 

- 

- 

29 1,030 21,408 

217 

4,384 20,501 26,790 45,470 60,562 59,168 23,667 33,512 35,138 31,913 26,556 26,044 1,624 

- 

- 

- 

- 

Recovery 60.0% 0.0% 0.0% 60.0% 60.0% 60.0% 60.0% 60.0% 60.0% 60.0% 60.0% 60.0% 60.0% 60.0% 60.0% 60.0% 60.0% 60.0% 60.0% 60.0% 60.0% 60.0% 0.0% 0.0% 0.0% 0.0% 


- 

- 

- 

- 

17 

618 12,845 

130 

2,630 12,300 16,074 27,282 36,337 35,501 14,200 20,107 21,083 19,148 15,933 15,626 

974 

- 

- 

- 

- 

Oxide Ore - Other Areas (kt) 52,537 

- 

- 

- 

- 

790 

538 1,040 3,149 

5,723 6,360 4,771 4,337 2,383 2,841 7,862 5,124 4,093 3,108 

418 

- 

- 

- 

- 

- 

- 



- 

- 

- 

- 

2,721 2,034 3,936 13,400 22,904 24,166 17,126 13,888 8,579 11,957 32,888 18,361 13,904 11,189 1,170 

- 

- 

- 

- 

- 

- 

Recovery 65.0% 0.0% 0.0% 65.0% 65.0% 65.0% 65.0% 65.0% 65.0% 65.0% 65.0% 65.0% 65.0% 65.0% 65.0% 65.0% 65.0% 65.0% 65.0% 65.0% 65.0% 65.0% 0.0% 0.0% 0.0% 0.0% 


- 

- 

- 

- 

1,769 1,322 2,559 8,710 14,888 15,708 11,132 9,027 5,576 7,772 21,377 11,934 9,038 7,273 

761 

- 

- 

- 

- 

- 

- 

Payable Metals 

Payable Copper (klbs) 747,689 

- 

- 

45,340 

50,113 

43,572 

43,745 

41,313 

40,166 

38,134 

39,018 

Income Statement ($000) 

Metal Prices 

Copper ($/lb.) $ 3.48 

$ 3.48 $ 3.48 $ 3.48 $ 3.48 $ 3.48 $ 3.48 $ 3.48 $ 3.48 $ 3.48 $ 3.48 $ 3.48 $ 3.48 $ 3.48 $ 3.48 $ 3.48 $ 3.48 $ 3.48 $ 3.48 $ 3.48 $ - $ - $ - $ - 

Revenues 

Copper Revenue ($ 000) $ 2,601,957 

$ 157,782 $ 174,392 $ 151,632 $ 152,232 $ 143,768 $ 139,776 $ 132,708 $ 135,783 $ 135,454 $ 137,373 $ 145,858 $ 160,153 $ 159,086 $ 143,657 $ 139,096 $ 130,013 $ 134,843 $ 124,689 $ 3,662 $ - $ - $ - $ - 

Total Revenues $ 2,601,957 $ - $ - $ 157,782 $ 174,392 $ 151,632 $ 152,232 $ 143,768 $ 139,776 $ 132,708 $ 135,783 $ 135,454 $ 137,373 $ 145,858 $ 160,153 $ 159,086 $ 143,657 $ 139,096 $ 130,013 $ 134,843 $ 124,689 $ 3,662 $ - $ - $ - $ - 

Operating Cost 

Mine Operations $ 743,021 $ - $ 3,390 $ 27,042 $ 25,517 $ 29,000 $ 29,400 $ 29,200 $ 41,100 $ 48,500 $ 49,200 $ 49,900 $ 47,800 $ 47,800 $ 45,644 $ 46,802 $ 49,200 $ 46,454 $ 41,912 $ 45,701 $ 37,134 $ 2,325 $ - $ - $ - $ - 

SXEW Plant $ 282,315 $ - $ - $ 16,871 $ 18,127 $ 16,406 $ 16,451 $ 15,811 $ 15,509 $ 14,974 $ 15,207 $ 13,419 $ 15,327 $ 15,969 $ 14,581 $ 16,969 $ 15,802 $ 15,458 $ 14,771 $ 15,136 $ 14,368 $ 1,159 $ - $ - $ - $ - 

Acid Cost $ 260,230 $ - $ - $ 14,330 $ 13,959 $ 14,082 $ 14,042 $ 14,120 $ 14,447 $ 14,847 $ 14,945 $ 14,699 $ 14,632 $ 14,329 $ 14,400 $ 15,178 $ 14,754 $ 14,594 $ 14,441 $ 14,024 $ 13,959 $ 449 $ - $ - $ - $ - 

General Administration $ 91,100 $ - $ - $ 4,992 $ 4,992 $ 4,992 $ 4,992 $ 4,992 $ 4,992 $ 4,992 $ 4,992 $ 4,992 $ 4,992 $ 4,992 $ 4,992 $ 4,992 $ 4,992 $ 4,992 $ 4,992 $ 4,992 $ 4,992 $ 1,248 $ - $ - $ - $ - 

Transportation $ 37,384 

- 

- 

2,267 2,506 2,179 2,187 2,066 2,008 

1,907 1,951 1,946 1,974 2,096 2,301 2,286 2,064 1,999 1,868 1,937 1,792 

53 

- 

- 

- 

- 

Total Operating Cost $ 1,414,051 

- 

3,390 

65,502 

65,100 

66,658 

67,072 

66,189 

78,057 

85,220 

86,295 

Property Tax $ - 

- 

- 

- 

- 

- 

- 

- 

- 

- 

- 

- 

- 

- 

- 

- 

- 

- 

- 

- 

- 

- 

- 

- 

- 

- 

Royalty - Arimetco $ 7,500 $ - $ - $ 2,773 $ 3,075 $ 1,652 $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - 

Royalty - 3rd Party $ 23,823 $ - $ - $ 2,386 $ 1,538 $ 1,330 $ 1,336 $ 1,259 $ 1,223 $ 1,158 $ 1,186 $ 1,201 $ 1,201 $ 1,278 $ 1,433 $ 1,398 $ 1,258 $ 1,216 $ 1,134 $ 1,178 $ 1,085 $ 25 $ - $ - $ - $ - 

Salvage Value $ (9,196) 

$ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ (9,196) $ - 

Reclamation & Closure $ 92,159 $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ 23,040 $ 23,040 $ 23,040 $ 23,040 $ - 

Total Production Cost $ 1,528,336 $ - $ 3,390 $ 70,661 $ 69,713 $ 69,640 $ 68,408 $ 67,448 $ 79,279 $ 86,378 $ 87,481 $ 86,157 $ 85,925 $ 86,463 $ 83,350 $ 87,626 $ 88,070 $ 84,712 $ 79,117 $ 82,968 $ 73,330 $ 28,297 $ 23,040 $ 23,040 $ 13,844 $ - 

Operating Income $ 1,073,620 $ - $ (3,390) $ 87,121 $ 104,679 $ 81,992 $ 83,823 $ 76,321 $ 60,497 $ 46,330 $ 48,301 $ 49,297 $ 51,448 $ 59,395 $ 76,802 $ 71,460 $ 55,587 $ 54,384 $ 50,896 $ 51,875 $ 51,360 $ (24,635) $ (23,040) $ (23,040) $ (13,844) $ - 

Initial Capital Depreciation $ 232,749 

$ 33,260 $ 57,000 $ 40,708 $ 29,070 $ 20,784 $ 20,761 $ 20,784 $ 10,381 $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - 

Mine Development $ - 

$ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - 

Sustaining Capital Depreciation $ 147,568 

$ 314 $ 910 $ 2,162 $ 2,685 $ 8,561 $ 13,532 $ 14,444 $ 17,271 $ 15,392 $ 14,082 $ 14,383 $ 11,993 $ 9,225 $ 7,673 $ 5,305 $ 3,725 $ 2,676 $ 1,484 $ 1,749 $ - $ - $ - $ - 

Total Depreciation $ 380,317 $ - $ - $ 33,574 $ 57,911 $ 42,870 $ 31,755 $ 29,345 $ 34,294 $ 35,229 $ 27,651 $ 15,392 $ 14,082 $ 14,383 $ 11,993 $ 9,225 $ 7,673 $ 5,305 $ 3,725 $ 2,676 $ 1,484 $ 1,749 $ - $ - $ - $ - 

Net Income After Depreciation $ 693,304 

- 

Income Taxes & Minerals Tax $ 179,087 

- 

Net Income After Taxes $ 514,217 

- 

(3,390) 

(3,390) 

53,546 

11,145 

42,401 

46,768 

9,821 

36,947 

M3-PN110127 

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20,000 

35,000 

38,923 

84,956 

20,000 

35,000 

39,475 

84,724 

20,000 

35,000 

41,913 

85,185 

17,403 

32,403 

46,021 

81,918 

$ 39,122 $ 52,068 $ 46,975 $ 26,204 $ 11,101 $ 20,650 $ 33,904 $ 37,366 $ 45,012 $ 64,809 $ 62,236 $ 47,914 $ 49,079 $ 47,171 $ 49,200 $ 49,875 $ (26,385) $ (23,040) $ (23,040) $ (13,844) $ - 

8,216 

30,906 

12,054 

40,014 

10,538 

36,438 

5,503 

20,701 

2,331 

8,770 

4,337 

16,314 

7,120 

26,785 

7,847 

29,519 

9,672 

35,340 

16,543 

48,266 

18,156 

33,156 

45,714 

86,227 

15,609 

46,627 

20,000 

35,000 

41,281 

86,812 

10,905 

37,009 

17,261 

32,261 

39,970 

83,496 

11,593 

37,486 

12,749 

27,749 

37,360 

77,983 

11,335 

35,836 

15,256 

30,256 

38,748 

81,790 

11,863 

37,337 

7,612 

22,612 

35,830 

72,244 

12,656 

37,219 

194 

676 

1,052 

5,233 

- 

(26,385) 

- 

- 

- 

- 

- 

(23,040) 

- 

- 

- 

- 

- 

(23,040) 

- 

- 

- 

- 

- 

(13,844) 

- 

- 

- 

- 

- 

-



Cash Flow 

Operating Income $ 1,073,620 $ - $ (3,390) $ 87,121 $ 104,679 $ 81,992 $ 83,823 $ 76,321 $ 60,497 $ 46,330 $ 48,301 $ 49,297 $ 51,448 $ 59,395 $ 76,802 $ 71,460 $ 55,587 $ 54,384 $ 50,896 $ 51,875 $ 51,360 $ (24,635) $ (23,040) $ (23,040) $ (13,844) $ - 

Working Capital 

Account Recievable (15 days) $ - $ - $ - $ (6,484) $ (683) $ 935 $ (25) $ 348 $ 164 $ 290 $ (126) $ 14 $ (79) $ (349) $ (587) $ 44 $ 634 $ 187 $ 373 $ (198) $ 417 $ 4,974 $ 150 $ - $ - $ - 

Accounts Payable (30 days) $ - $ - $ 279 $ 5,105 $ (33) $ 128 $ 34 $ (73) $ 975 $ 589 $ 88 $ (110) $ (19) $ 38 $ (269) $ 354 $ 48 $ (273) $ (453) $ 313 $ (785) $ (5,508) $ (430) $ - $ - $ - 

Inventory - Parts, Supplies $ - $ - $ (440) $ (660) $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ 1,100 $ - $ - $ - $ - 

Total Working Capital $ - $ - $ (161) $ (2,039) $ (716) $ 1,063 $ 9 $ 275 $ 1,139 $ 879 $ (38) $ (97) $ (98) $ (311) $ (856) $ 398 $ 682 $ (85) $ (80) $ 114 $ (367) $ 566 $ (280) $ - $ - $ - 

Capital Expenditures 

Initial Capital 

Mine $ 48,000 $ 2,400 $ 43,200 $ 2,400 $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - 

SXEW Plant $ 114,310 $ 5,716 $ 102,879 $ 5,716 $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - 

Acid Plant $ 65,439 $ 3,272 $ 58,895 $ 3,272 $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - 

Owners Cost $ 5,000 $ 250 $ 4,500 $ 250 $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - 

Land Acquisition $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - 

Mine Development $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - 

Sustaining Capital 

Mine $ 83,600 $ - $ - $ 2,200 $ 2,600 $ 2,600 $ - $ 15,000 $ 5,000 $ 18,800 $ 19,400 $ 1,100 $ 8,100 $ 8,800 $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - 

SXEW Plant $ 63,968 $ - $ - $ - $ - $ 5,383 $ - $ 31,489 $ 50 $ 8,740 $ 151 $ 135 $ 5,662 $ 250 $ 3,420 $ 3,825 $ 4,862 $ - $ - $ - $ - $ - $ - $ - $ - $ - 

Total Capital Expenditures $ 380,317 $ 11,637 $ 209,474 $ 13,837 $ 2,600 $ 7,983 $ - $ 46,489 $ 5,050 $ 27,540 $ 19,551 $ 1,235 $ 13,762 $ 9,050 $ 3,420 $ 3,825 $ 4,862 $ - $ - $ - $ - $ - $ - $ - $ - $ - 

Cash Flow before Taxes $ 693,304 $ (11,637) $ (213,026) $ 71,244 $ 101,363 $ 75,072 $ 83,833 $ 30,107 $ 56,587 $ 19,669 $ 28,712 $ 47,965 $ 37,588 $ 50,034 $ 72,526 $ 68,034 $ 51,407 $ 54,299 $ 50,816 $ 51,990 $ 50,992 $ (24,069) $ (23,319) $ (23,040) $ (13,844) $ - 

Cummulative Cash Flow before Taxes $ (11,637) $ (224,663) $ (153,419) $ (52,056) $ 23,017 $ 106,849 $ 136,956 $ 193,543 $ 213,212 $ 241,924 $ 289,889 $ 327,478 $ 377,512 $ 450,038 $ 518,072 $ 569,478 $ 623,777 $ 674,594 $ 726,583 $ 777,576 $ 753,506 $ 730,187 $ 707,147 $ 693,304 $ 693,304 

1.0 

1.0 

0.7 

- 

- 

- 

- 

- 

- 

- 

- 

- 

- 

- 

Taxes 

Income Taxes $ 179,087 $ - $ - $ 11,145 $ 9,821 $ 8,216 $ 12,054 $ 10,538 $ 5,503 $ 2,331 $ 4,337 $ 7,120 $ 7,847 $ 9,672 $ 16,543 $ 15,609 $ 10,905 $ 11,593 $ 11,335 $ 11,863 $ 12,656 $ - $ - $ - $ - $ - 

Cash Flow after Taxes $ 514,217 $ (11,637) $ (213,026) $ 60,099 $ 91,542 $ 66,857 $ 71,778 $ 19,569 $ 51,084 $ 17,337 $ 24,376 $ 40,845 $ 29,741 $ 40,362 $ 55,982 $ 52,425 $ 40,502 $ 42,706 $ 39,481 $ 40,127 $ 38,337 $ (24,069) $ (23,319) $ (23,040) $ (13,844) $ - 

Cummulative Cash Flow after Taxes $ (11,637) $ (224,663) $ (164,564) $ (73,022) $ (6,165) $ 65,613 $ 85,182 $ 136,266 $ 153,604 $ 177,979 $ 218,825 $ 248,566 $ 288,928 $ 344,911 $ 397,336 $ 437,838 $ 480,544 $ 520,025 $ 560,152 $ 598,489 $ 574,419 $ 551,100 $ 528,060 $ 514,217 $ 514,217 

1.0 

1.0 

1.0 

0.1 

- 

- 

- 

- 

- 

- 

- 

- 

- 

- 

- 

- 

- 

- 

- 

- 

- 

- 

- 

Economic Indicators before Taxes 

NPV @ 0% 0% $ 693,304 

NPV @ 5% 5% $ 395,451 

NPV @ 8% 8% $ 284,138 

IRR 29.3% 

Payback Years 2.7 

Economic Indicators after Taxes 

NPV @ 0% 0% $ 514,217 

NPV @ 5% 5% $ 288,066 

NPV @ 8% 8% $ 201,576 

IRR 24.2% 

Payback Years 3.1 

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23 ADJACENT PROPERTIES 

23.1 SINGATSE PEAK SERVICES PROPERTIES 

During April 2011, Singatse Peak Services, LLC (SPS), a wholly owned subsidiary of Quaterra 

Alaska Inc., purchased the historic Yerington Mine Copper Property comprising over 12,000 

acres of private lands and unpatented lode mining claims south of and contiguous with Quaterra 

Alaska’s MacArthur property Figure 23-1. The Yerington Mine Copper Property was operated 

from 1952 to 1977 by The Anaconda Company and from 1977 to 1979 by Atlantic Richfield 

Corporation. 

During February 2012, SPS published a NI 43-101 compliant independent resource estimate at 

the Yerington Mine Copper Property. As cited in their Technical Report, resources at a 0.2% Cu 

cutoff are shown in Table 23-1. 

Table 23-1: Singatse Peak Services, LLC – Yerington Mine Resources, Feb. 2012 

Grade 

Cutoff 

% Cu 

Tons 

(x1000) 

Average Grade 

% Cu 

Contained Copper 

(lbs x 1000) 

Measured and Indicated 


Material 

0.2 9,445 0.3 57,237 


Inferred 

0.2 71,781 0.3 429,968 


Material 

0.2 8,596 0.28 47,347 

Primary Material 0.2 63,918 0.25 322,530 

SPS’s assets on the Yerington Mine Copper Property also include over 120 million tons of 

resource piles (“Residuals”) representing sub-grade stripping material from the Yerington mine 

vat leach tailings (VLT) representing the oxide tailings from Anaconda’s copper oxide vat 

leaching, and historic heap leach pads previously mined by Arimetco. Approximately 44 million 

tons consisting primarily of sub-grade material (Stockpiles W-3 and S-23 in Table 23-2) and 

VLT were heaped and leached (including an estimated 6 million tons of oxide material from the 

MacArthur mine) by Arimetco International Inc. during 1989-1998. 

No copper extraction from the Arimetco heaps or mining has occurred since Arimetco closure in 

1999, leaving an estimated (non-compliant NI43-101) over 300 million pounds of contained 

copper in the residuals as summarized in Table 23-2. References 2 through 5 shown on the table 

refer to documents published by the USEPA (EPA), as listed in Section 27, References. Work is 

ongoing by SPS to further characterize the Yerington residuals, including extensive drilling and 

metallurgical testing to assess the viability of reprocessing some or all of these materials. 

Depending on the results of the ongoing work, some of these materials may be integrated with 

the MacArthur Oxide project in the future. The residuals are further discussed in Section 25 of 

this TR. 

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Table 23-2: Yerington Mine Residual Copper Resources, SRK March, 2012 (Non NI43-101 

Compliant) 

Material Type Ore (tons) 

Volume (cu. 

yds.) 

Total Cu 

Grade (%) 

Contained Cu 

Expected Leach 

Recovery 

Oxide Tails (VLT) 1,3,5 57,571,505 35,545,074 0.130 149,686 75% 

Oxide Low-Grade W-3 1,4 19,643,073 12,127,779 0.200 78,572 60% 

Sulfide Low-Grade 4, 2,316,440 1,430,187 0.200 9,266 85% 

Phase 1/2 HLP 1,2 2,104,570 1,362,710 0.099 4,159 50% 

Phase 3 HLP 4 1,2 8,547,269 5,147,407 0.120 20,513 50% 

Phase 3 HLP S 1,2 10,117,573 5,836,837 0.083 16,714 50% 

Phase 4 Slot HLP 1,2 12,927,862 8,793,567 0.091 23,399 50% 

Phase 4 HLP 1,2 11,556,016 6,539,352 0.075 17,242 50% 

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(k lb) 

Total 124,784,308 76,782,913 0.128 319,551 

Notes: 

1 Volume based on: SRK 2010 digitization and volume calculations using MineSight 3D Software. 

2 Density based on: Draft Supplemental RI Report_OCT_2010 - Page 47. 

3 Grade based on: AnacondaArimetco_RI_Report.pdf - Page 170-172. 

4 Grade based on: VLT XRF DSR July 2010 - Page 99. 

5 Grade based on: HistoricalSummaryReport-YeringtonMine-2010-10.pdf - Page 19. 

23.2 OTHER PROPERTIES 

The following information is presented as an indication of the types and magnitude of similar 

surrounding deposits and mines. The deposits presented are all within a few miles of the 

MacArthur Project and have mineralization that is similar in nature to the MacArthur Project. 

The Ann Mason resources, based upon a 0.3% Cu cutoff, are taken from the March 2012 NI 43- 

101 Technical Report completed for Entree Gold. The Bear-MacArthur-Lagomarsino resources 

referenced (which were obtained from MineMarket.com in 2004) have not been classified 

according to current CIM standards. A portion of the Bear-MacArthur-Lagomarsino prospect 

underlines the Yerington Site. 

Table 23-3 lists historic resource estimates for two porphyry copper deposits in the Yerington 

district.



Table 23-3: Adjacent Property Resource Estimates 

Adjacent Property Resource Estimates 

May 2012 

Adjacent Property Name Ore Tons Average Grade Contained Cu 

(kTons) (% Cu) (kTons) 

Contained Cu 

(000s lbs) 

Ann Mason Deposit 1 1,084,000 0.37 4,460 8,920,000 

Bear-MacArthur- 

Lagomarsino Deposit 

500,000 0.40 2,000 4,000,000 

Total all deposits 1,584,000 0.37 6,460 12,920,000 

1 Sum of Indicated and Inferred Resources taken 

from March 2012 NI 43-101 Technical Report and 

Updated Mineral Resource Estimate 

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Figure 23-1: Adjacent Properties 

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24 OTHER RELEVANT DATA AND INFORMATION 

24.1 RE-PROCESSING OF YERINGTON RESIDUALS 

24.1.1 Introduction 

A scoping study was prepared by SRK Consulting (U.S.), Inc. (SRK) for Singatse Peak Services 

LLC (SPS), a wholly-owned subsidiary of Quaterra Resources Inc., to evaluate the technical and 

economic feasibility for re-processing residual spent ore, waste rock, and oxide leach tailings at 

the now closed Yerington Copper Mine located near the town of Yerington in Lyon County, 

Nevada. The Yerington Copper Mine is owned by Singatse Peak Services (SPS), a subsidiary 

company of Quaterra Resources, and is located within four miles of the MacArthur Pit with 

possibility for sharing infrastructure facilities. 

The scoping study, dated March 12, 2012, was intended for SPS internal use only and does not 

conform to CIM NI 43-101 standards for public disclosure. Results of the scoping study are 

presented here for the purpose of highlighting potential synergies between the MacArthur 

Copper Project and the re-processing of residual ore stockpiles and tailings at the Yerington 

Copper Mine and to get an early look at the potential economic benefit of a combined project. 

24.1.2 Residual Copper Resources 

SRK quantified four material types at Yerington as potential resources for re-processing and 

extracting residual copper at Yerington. Three of the four mineral types are oxide in nature and 

considered suitable for combining with the MacArthur process facilities. The fourth material is a 

low grade sulfide more amenable to a mill / flotation circuit. The three oxide material types 

considered for a combined project with MacArthur are noted below. 

a) Crushed vat leach tailings (VLT) from the former Anaconda processing of oxide ore. 

b) A low grade run-of-mine (ROM) oxide stockpile (W-3) from the Yerington pit that was 

below Anaconda’s cutoff grade of 0.3% copper for copper ore. 

c) Five heap leach pads (HLPs) built and operated by Arimetco containing mostly W-3 

oxide, VLT, and small additions of copper oxide from the MacArthur mine. 

24.1.2.1 Vat Leach Tailings (VLT) 

The VLT stockpile area covers approximately 500 acres primarily on private land owned by SPS 

with an average height of approximately 100 feet. The tops surfaces are composed of multiple 

benches and VLT mounds channeled to prevent storm water runoff. The VLT resource was 

estimated to be approximately 57.6 million tons of ore at an average grade of 0.13% Cu. The 

resource estimate was based on volumetric calculations and density measurements from 

historical data and recent test work. The grade determination was based on averages of samples 

reported in a recent VLT characterization study by Atlantic Richfield (ARC, 2010). The average 

grade of the VLT material from this report was 0.13% Cu. Subsequently, METCON (2011) 

conducted head assays for materials used in six column leach tests, which averaged 0.18% Cu. 

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SRK believed that the column test average grade was biased high by one sample with a head 

assay of 0.356% Cu; therefore SRK used the 0.13 % Cu assay from the Atlantic Richfield study. 

SPS contracted METCON Research of Tucson, Arizona to conduct column leach tests on six 

samples of VLT stockpile material with head grades ranging from 0.35% down to 0.06%. The 

column tests were run in locked cycle for 93 days followed by an eight day rinse and drain down. 

Testing indicated that copper recovery exceeded 70% on all but two samples with low head 

grades and low (~50%) acid soluble Cu to total Cu ratios. Test work also indicated a gangue acid 

consumption of approximately 30 lb. acid per ton of ore in 90 day column leach tests. Gangue 

acid consumption plots show that consumption continues at a constant rate over the 93 day leach 

cycle indicating that a shorter leach cycle should produce high copper recoveries at a reduced 

acid consumption. The design basis for the VLT material was 18 lbs. acid per ton ore based on 

20 day leach results from METCON (METCON, 2011). 

24.1.2.2 W-3 Low Grade Oxide 

Anaconda originally stockpiled low grade oxide that was below their operating cutoff of 0.3% 

Cu, but above their 0.2% threshold. The stockpile is north of the Yerington open pit and is 

primarily on land controlled by BLM. The current W-3 low grade oxide stockpile covers 

approximately 80 acres, with a maximum height of 210 feet, and averaging about 160 feet. Side 

slopes are generally 1.4H: 1V. 

The W-3 resource is estimated to be approximately 19.6 million tons at an average grade of 0.2% 

Cu. The resource estimate was based on volumetric calculations and density measurements from 

historical data and recent test work. The acid consumption for the W-3 material was set at 35 lbs. 

acid per ton of ore. 

24.1.2.3 Heap Leach Pads (Arimetco) 

Arimetco constructed five distinct heap leach pads (HLPs), built in four phases, covering nearly 

250 acres during their operation between 1990 and 1999 (see Figure 24-1). Phase 1 is located 

immediately north of the Yerington open pit and southeast of the original SX/EW facility. Phase 

II is contiguous with phase I, extending it to the northwest. Phase III consists of two separate 

lined heap leach pads, with Phase III South and Phase III 4X both located north of the access 

road and west of the historic process areas. Phase IV also consists of two separate HLPs; 1) the 

Slot bordering the eastern property boundary and including portions of Anaconda’s W-3 

stockpile; and 2) the VLT heap located northeast of the VLT footprint. It should be noted that 

much of the Phase III 4X and Phase IV Slot HLPs reside on BLM administered land. 

The Arimetco heap leach pads consist of mostly minus 6-inch material sourced from the W-3 

oxide stockpile, with some MacArthur ore, stacked approximately 100 to 120 feet high in 

nominal 20-foot high lifts. The exception is the Phase IV VLT, which is comprised of primarily 

VLT material. The estimated resource for all the heap leach pads is approximately 45.3 million 

tons at an average grade of approximately 0.1% Cu, representing the post-leaching residual grade 

of the ore that carried between 0.2% and 0.3% Cu when originally stacked. The grades were 

discounted based on Arimetco production reports and recent drilling results (CH2M HILL, 

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2008). Acid consumption should be similar to the VLT material, even after partial leaching, and 

has been set at 30 lbs. acid per ton of ore for the SRK scoping study. 

The total resource identified in the scoping study is shown in Table 24-1 below. For the scoping 

study, SRK used 75% recovery for the VLT, 60% recovery for the W-3 and 50% for the 

Arimetco HLPs. The recovery estimates were based on historic records of production from the 

Anaconda vat leach operation (70%), and the Arimetco recovery records from the initial leach of 

the HLP materials in phase I to IV (53%) (Sawyer, 1999). Recent column leach test work on 

both VLT and MacArthur (METCON, 2011) established a basis for slightly higher forecasts for 

the VLT. 

Table 24-1: Yerington Residual Oxide Copper Resources, SRK March 2012 

Average 

Total Cu 

Grade 

Expected 

Leach 

Recovery 

Contained 

Extracted Cu 

Material Type Ore (tons) 

Cu (lb.) 

(lb.) 

Oxide Tails (VLT) 57,572,000 0.130% 149,686,000 75% 112,265,000 

Oxide Low Grade (W-3) 19,643,000 0.200% 78,572,000 60% 47,143,000 

Phase I and II HLP 2,105,000 0.099% 4,159,000 50% 2,080,000 

Phase III HLP 4X 8,547,000 0.120% 20,513,000 50% 10,257,000 

Phase III HLP S 10,118,000 0.083% 16,714,000 50% 8,357,000 

Phase IV Slot HLP 12,928,000 0.091% 23,399,000 50% 11,700,000 

Phase IV HLP 11,556,000 0.075% 17,242,000 50% 8,621,000 

Totals 122,469,000 0.127% 310,285,000 65% 200,423,000 

24.1.3 Mining Methods 

24.1.3.1 Vat Leach Tailings (VLT) 

SRK proposed an on/off leach pad for the VLT material in order to reduce acid consumption. 

The VLT ore would be mined from the existing stockpile, agglomerated, and conveyed to one of 

four leach cells comprising the on/off leach pad. Ore will be stacked to a height of 24 feet for 

leaching. After the leach cycle, the pad would be rinsed and drained before removing the spent 

leached material from the pad and stored in a lined VLT storage facility. Removal of the leached 

material would be by conventional mining equipment, including rubber tired loaders and 100 ton 

off-highway trucks. The liner system for the on/off pad would tie to the existing VLT stockpile 

liner and the new spent VLT lined storage facility. The entire operation, therefore, would be 

carried out on containment. 

The mining rate for the VLT material was set at 7.9 million tons per year with a life of mine of 

7.25 years. The daily mining rate was estimated to be approximately 21,700 tons per day. The 

leach solution flow from the VLT leach pad was set at 3,000 gpm based on an irrigation rate of 

0.04 gpm / ft. 2 . The solution grade was estimated to be 1.18 gpl Cu. 

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24.1.3.2 W-3 Low Grade Oxide 

The W-3 material will be leached by a conventional heap leach pad. The leach pad would be 

constructed as a continuation of the “Slot Pad” concept initiated by Arimetco in the late 1990’s. 

The remaining material in the existing slot (Slot 1) would be removed with loader and trucks and 

slot 1 will be lined. W-3 ore from the next slot (Slot 2) would be mined and placed in the new 

Slot 1 pad and Slot 2 would be lined. The sequence would continue until the entire W-3 stockpile 

has been placed on a lined leach pad and leached. It is anticipated that four phases of slot 

expansions would be lined to create a total contiguous lined area of 2.70 million square feet. 

The mining rate for the W-3 material was set at 2.7 million tons per year with the same life of 

mine of 7.25 years. The daily mining rate is estimated to be approximately 7,400 tons per day. 

The leach solution flow from the W-3 leach pad also was set at 3,000 gpm with an expected 

grade of 0.5 gpl Cu. 

24.1.3.3 Heap Leach Pads (HLP) 

For the scoping study, it was assumed that the existing heap leach pads from the Arimetco 

operation would be leached again in place. The horizontal top surfaces of these pads would be 

ripped with dozers to enhance the permeability at the surface. Some repairs are expected to be 

required to the existing perimeter ditches and ponds prior to re-leaching. In some cases the 

existing ponds may need to be replaced. The existing heap leach pads would be leached until the 

cost of power and reagents is greater than the revenue generated from the copper production. The 

pads would then drain down, be re-graded to a 3H: 1V slope and capped. 

For the purpose of the scoping study, SRK assumed the time of leaching would coincide with the 

completion of the VLT and W-3 leaching, or 7.25 years. The heap leach pads would be leached 

in sequence to minimize the capital expenditures for pumps, piping and the SX/EW plant. A 

maximum of two heap leach pads would be under leach at any one time. The PLS flow from the 

HLPs was set at 3,000 gpm with an expected grade of 0.4 gpl Cu. 

For the purposes of the scoping study, it was assumed that the HPL leach solution would be 

staged in series with the W-3 leach pads to produce a combined PLS flow of 3,000 gpm at a 

grade of 0.9 gpl Cu. The total PLS flow from the Yerington leach operation is expected to be 

6,000 gpm at a grade of 1.05 gpl. This total PLS flow would be combined with the MacArthur 

leach solution (10,400 gpm at 1.0 gpl Cu) providing a total PLS flow to the MacArthur SX/EW 

plant of 16,400 gpm at about 1.0 gpl. 

24.1.4 Capital Cost Summary 

SRK developed capital and operating costs for three major case scenarios. The base case (Case 

1) included leaching the VLT, W-3 stockpile and re-leaching the Arimetco heap leach pads. 

Case 2 assumed that the VLT and W-3 stockpile would be processed. Case 3 assumed only the 

VLT material would be leached. Capital cost estimates were modified in each case to reflect the 

change in assumptions. A second set of cases (Case 1A, Case 2A, and Case 3A) were also run 

without the SX/EW facilities, assuming the leach solution from the residual copper leach 

operation at Yerington would be combined and treated in the MacArthur SX/EW facility. 

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M3 updated the capital cost and operating cost for the SX/EW and sulfuric acid plant in order to 

accommodate the increased solution flow and increased sulfuric acid consumption in the 

MacArthur facilities to accommodate the SRK Case 1A. The combined capital cost for the 

Yerington residual leach operation and the increased MacArthur operation are summarized in 

Table 24-2 below. 

Table 24-2: Combined Yerington Oxide Residuals / MacArthur Mine Capital & 

Sustaining Costs 

Initial Capital 

Sustaining 

Capital 

MacArthur Copper Operation 

Mine Equipment $48,000,000 $83,600,000 

EX/EW $138,737,000 $63,968,000 

Sulfuric Acid Plant $100,105,000 $0 

Owner's Cost $5,000,000 

Sub Total $291,842,000 $147,568,000 

Yerington Risidual Leach Operation 

Mine Equipment $28,694,000 $1,306,000 

Process & Leach Pads $29,724,000 $18,462,000 

Infrastructure $1,300,000 $0 

Owner's Cost $408,000 $2,792,000 

Sub Total $60,126,000 $22,560,000 

Total Combined Yerington/MacArthur $351,968,000 $170,128,000 

24.1.5 Operating Costs 

Operating costs for the combined Yerington leach operation and MacArthur leach operation is 

summarized in Table 24-3 below. The SKR operating cost was adjusted to account for the cost of 

sulfuric acid from an onsite sulfuric acid plant instead of purchased acid. The MacArthur sulfuric 

acid cost will also see a reduction because of the economy of scale with a larger acid plant. 

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Table 24-3: Combined Yerington Oxide Residuals / MacArthur Mine Operating Costs 

Annual Cost $ / lb. Cu 

MacArthur Copper Production, Lbs. 41,538,000 

MacArthur Copper Operation 

Mine $41,279,000 $1.00 

SX/EW $19,903,000 $0.48 

Acid $12,633,000 $0.31 

G&A $5,082,000 $0.12 

Transportation $2,077,000 $0.05 

Sub-Total $80,974,000 $1.96 

Yerington Copper Production, Lbs. 27,666,000 

Yerington Residual Leach Operation 

Mine $11,204,000 $0.40 

Heap Leach $12,925,000 $0.47 

Solution Pumping $7,416,000 $0.27 

G&A $291,000 $0.01 

Transportation $1,383,000 $0.05 

Sub-Total $33,219,000 $1.20 

Combined Yerington / MacArthur $114,193,000 $1.65 

SRK based their capital cost estimate on their in-house experience with similar projects, scaled 

to the size of this project. Costs for many of the equipment items used were from recent vendor 

quotes. Where recent cost data was not available, commercially available mining cost services, 

such as InfoMine, was used. Mine equipment was selected with a life cycle of 30,000 to 40,000 

hours, which equates to a 7 to 8 year mining operation. SRK considered using the same 

equipment for haulage at both the VLT off-loads and to move the W-3 stockpile. SRK assumed 

that the existing buildings on site would be used for office and warehousing. A new truck shop 

was provided to support the proposed fleet of 100 ton trucks. SRK applied a contingency of 30% 

on all their estimates. 

M3 factored the cost of the SX/EW facility and sulfuric acid plant from the PEA capital cost 

estimate to the new capacities. The capital cost of the sulfuric acid plant was factored from 640 

tons per day plant to 1,220 tons per day. The solvent extraction plant, originally sized for 

MacArthur, can be used in combination with the Yerington residual PLS by changing the 

configuration of the 3 extraction settlers from a 2-series, 1-parallel configuration to 3 parallel 

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settlers. The capital cost of the tank farm, electrowinning, main substation, and reagent areas 

were factored based on the increase in copper production. 

24.1.6 Economic Analysis 

A financial analysis was prepared for the combined MacArthur and Yerington operation using 

the same parameters as for the stand-alone MacArthur Project and the preliminary Yerington 

oxide residuals case without a SX/EW facility. The combined life of mine recovered copper is 

948,266,000 pounds with combined revenues of $3.3 billion at the $3.48 per pound price of 

copper. The Net Present Value (NPV), Internal Rate of Return (IRR) and payback period before 

and after taxes are shown in Table 24-4 below. 

Table 24-4: Combined Yerington Oxide Residuals / MacArthur Mine Economic Indicators 

Economic Indicators Before Taxes After Taxes 

$000 $000 

NPV at 8% Disount Rate $435,681 $308,307 

IRR, % 32.9% 26.5% 

Payback, years 2.5 2.9 

The preliminary economic evaluation for the combined MacArthur heap leach operation and the 

Yerington residual re-processing operation is estimated to add over $100 million to the standalone 

MacArthur NPV at 8% discount rate, increase the IRR by over 2 percentage points, and 

reduce the payback period by approximately 2 months. 

Although the Yerington resource determination and recoveries are not sufficiently defined to 

comply with NI 43-101 reporting standards, there is sufficient justification for further 

investigation of the combined MacArthur heap leach operation and re-processing of Yerington 

residual materials. This work is ongoing and would be included with the pre-feasibility work for 

a combined oxide project. See Figure 24-1. 

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Figure 24-1: Yerington Mine Residuals 

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25 INTERPRETATION AND CONCLUSIONS 

The intent of this report is to incorporate previous resource information and technical reports 

prepared earlier with additional resource information and the Preliminary Economic Assessment 

(PEA). The results of this PEA suggest that the project may be technically feasible utilizing 

ROM heap leaching and solvent extraction / electrowinning technology and may be 

economically viable based on the resources, grade, and recovery information presented to date. 

There is potential to further enhance the project economics by integrating the MacArthur Copper 

Project with other copper oxide resources owned by Quaterra Resources, Inc. and Singatse Peak 

Services (SPS) in the Yerington District. Further work will be necessary, however, to quantify 

the other resources grade and recoveries to comply with NI 43-101 standards of disclosure for 

mineral projects. 

It is noted that a PEA should not be considered to be a pre-feasibility study or feasibility study, 

as the technical and economic viability of the project has not been demonstrated at this time. A 

PEA is preliminary in nature and includes Inferred Mineral Resources that are considered to be 

too geologically speculative at this time to have the economic considerations applied to them to 

be categorized as Mineral Reserves. 

25.1 RESOURCES 

The measured and indicated oxide and chalcocite resource for the project, at a cutoff grade of 

0.12% total copper is 159 million tons containing 675.5 million pounds of copper. The inferred 

oxide and chalcocite resource is 243.4 million tons at a cutoff grade of 0.12% containing 979.5 

million pounds of copper. 

The primary sulfide measured and indicated resource at a 0.15% total copper cutoff is 1.1 million 

tons containing 6.4 million pounds of copper. The inferred sulfide resource at the 0.15% cutoff is 

134.9 million tons containing 764 million pounds of copper. 

Further exploration drilling may enhance the project resources and reduce project risk. 

25.2 MINING METHODS 

Mining at the MacArthur pit will be by open pit at an approximate rate of 41,000 tons per day 

over an 18 year life of mine. Approximately 271 million tons of ore will be mined over the life 

of mine with an average waste to ore ratio of 0.90. There are three final pits consisting of the 

main MacArthur pit, the North pit and Gallagher pit. Mining will start in the main MacArthur pit 

and progress to the North pit, then to Gallagher pit and ending at the outer limits of the main 

MacArthur pit in the last phase. 

Waste dumps are located to the north and to the south of the pits with some pit backfill in the 

North pit area extending to the west of the North pit. 

There were no geotechnical studies for the pit slope angles performed for this PEA. No extensive 

condemnation drilling has been done in the north and south waste dump areas. Further 

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optimization of the mine plan, including waste rock management, will be completed in the next 

phase of the project. 

25.3 METALLURGY 

During the PEA review of historic and recent metallurgical test work for the MacArthur Project, 

several issues were identified that require additional test work to improve understanding of these 

issues and what impact they may have on the project. This work will be undertaken as part of the 

PFS which will include a comprehensive Phase II metallurgical test program. 

25.3.1 Run-of-Mine Heap Leaching 

The PEA was based on ROM ore truck dumping to a permanent heap leach pad. Based on the 

project ore grade of 0.211% total copper and high projected copper extraction, this approach 

provides simplicity of ore processing while optimizing acid consumption. 

Historical test work provides limited ROM data for copper extraction and acid consumption. 

However, MacArthur ROM ore was successfully processed by Arimetco with good copper 

extraction and acid consumption, supporting the ROM leaching approach. The proposed Phase 

II PFS metallurgical program will address particle size vs. copper extraction and further evaluate 

acid consumption. 

During the PFS, the Phase II test work program will be performed to provide data from which to 

make a final determination on optimization of leach particle size. During the Phase II test work 

program, a number of sites will be selected from the present bench faces in the MacArthur Pit. 

Using a portable screen and loader, this material will be screened to provide data on ROM top 

particle size distribution. 

25.3.2 Spatial Variability of In-Situ Size Distribution 

The 2011 column leach study by METCON Research included 32 column tests on PQ size core 

from 32 separate drill holes which spatially provided preliminary representivity of the block 

model. This study showed significant variation in particle size distribution from hole to hole. In 

heap leaching, particle size distribution is extremely important, particularly in particle sizes less 

than 100 mesh, and more importantly, less than 200 mesh. During the PFS, this issue will be well 

defined. 

25.3.3 Chemical Degradation of the Ore during Leaching 

The 2011 METCON Research column study also identified chemical degradation in some 

columns during leaching, impacting post leach size distribution. It is likely that the ore was over 

acidified during leaching resulting in un-necessary chemical degradation. However, this is likely 

not to be the only cause of variability in size distribution from drill hole to drill hole. This issue 

is extremely important to understand and will be studied during pre-feasibility metallurgical 

testing. 

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25.3.4 Permeability and Agglomeration 

Considering both the in-situ size distribution variability and chemical degradation during 

leaching, permeability of leached ore must be better understood as part of future design efforts. 

To accommodate uncertainty at this stage of project definition, the PEA design includes one 

inter-lift liner to be placed at approximately half the final pad elevation. 

Pre-feasibility test work will address spatial variations in size distribution and chemical 

degradation using size distribution measurements of column head and tails and analytical results. 

Column tails will be subjected to permeability/consolidation testing to define the ultimate 

permeability of multi-lift permanent leach pad processing. 

25.3.5 Spatial Variability of Copper Extraction and Acid Consumption 

The METCON column study also identified spatial variability in copper extraction and acid 

consumption. Moving away from the existing MacArthur Pit to the periphery zones of the block 

model tends to show reducing copper extraction and increasing acid consumption. 

The relationship of copper extraction and acid consumption versus depth in the deposit is also 

not well understood from existing test work and must be better quantified. 

The mine plan as provided by Independent Mining Consultants, Inc. (IMC) in this PEA has 

defined the materials to be extracted during the LOM. The PFS will aim to determine optimum 

processing techniques addressing both the spatial variation in copper extraction and acid 

consumption. Crushing and/or agglomeration, if justified, would distribute acid more efficiently 

through the ore resulting in reduced acid consumption and increased copper extraction, but at 

higher operating cost. This trade off will be further evaluated in future design studies. 

25.3.6 Relationship of Total Iron Mineralization to Acid Consumption 

One relationship that is clear in leaching MacArthur Project ore is the relationship between iron 

mineralization and acid consumption. Acid consumption in bottle roll and column testing 

consistently shows a linear increase in acid consumption versus time. Acid consumption tends to 

mirror iron extraction. This relationship will require optimal control of available free acid during 

leaching to maximize copper extraction while minimizing iron extraction. 

Again, the pre-feasibility Phase II metallurgical program will be designed to study this 

relationship including mineralogical studies. XRD, XRF, or QEMSCAN studies will provide 

total iron content and distribution of the iron species to determine which iron minerals are largely 

responsible for acid consumption. 

25.4 ECONOMIC ASSESSMENT 

Based on the historical three year average price of copper of $3.48 per pound, M3 engineering 

and Technology Corporation concluded that the after tax preliminary economic assessment of the 

project would provide a net present value, at an 8% discount rate, of approximately $201.5 

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million with an internal rate of return of 24.2 % and a payback period of 3.1 years. The 

preliminary economic assessment was based on an onsite sulfuric acid plant with a byproduct 

electrical power credit, an initial capital investment of $232.7 million and sustaining capital of 

$230.5 million over an 18 year life of mine. The overall average life of mine operating costs was 

calculated to be approximately $1.89 per pound of recovered copper. 

M3 concluded that the economic indicators for the stand alone MacArthur Copper Project has a 

potential for development as a large-scale copper oxide heap leach operation. There are also 

opportunities for enhanced economic indicators by combining the MacArthur Copper Project 

with the re-processing of residual vat leach tails, waste rock and historic leach pads at the 

existing Yerington site. It is noted that the resources for the residual material at the Yerington 

site are not NI 43-101 compliant at this time and will require further work to quantify the 

resources. 

25.5 RISKS 

The project risks identified at this time are listed below. Using a staged approach to advance the 

project to full production will allow Quaterra Resources, Inc. to adequately assess the risks and 

associated costs and develop mitigation strategies before progressing to the next stage. 

a) Further drilling may identify increases or decrease in resource tonnage and ore grade. 

b) Further metallurgical testing may demonstrate variable recoveries and acid consumption, 

resulting in changes to the economic indicators for the project. 

c) The cost of sulfur delivered to site for the manufacture of sulfuric acid may be higher or 

lower than estimated in the PEA resulting in changes to operating costs for the project. 

d) The capital cost estimates for initial capital and sustaining capital may be higher or lower 

than estimated in this PEA resulting in changes to the economic indicators for the project. 

e) The future price of copper may fall below the level necessary to sustain a viable 

operation. Likewise, the copper price may increase above the base resulting in better 

economic indicators for the project. 

f) The heap leach pad lift heights and overall height may impact the permeability of the 

leach solution through the heap, changing overall copper recovery. 

g) More or less fines in the run-of-mine ore may change the permeability of the leach 

solution, affecting overall copper recovery. Agglomeration of the ore may be required if 

fines are excessive. 

h) The allowances for refurbishing existing buildings at the existing Yerington site may not 

be sufficient, increasing the capital cost for the Project. 

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26 RECOMMENDATIONS 

This section summarizes the recommendations made for Quaterra Resources, Inc. consideration 

as the project progresses to the next phase. 

26.1 METALLURGY TEST PROGRAM 

Additional drilling and metallurgical test work is required to complete a PFS or FS. Based on 

the mine plan prepared for the PEA, LOM tonnage and grade has been generated along with 

tonnage and grade mined per year. This plan also includes annual mined material segregated by 

MacArthur oxide ore, oxide ore from areas other than MacArthur Pit, and tonnage and grade of 

mixed oxide/secondary sulfide ores. 

With the mine plan and phasing complete, a PFS or FS metallurgical drilling program can be 

defined. Geology and mine planning personnel in consultation with metallurgical personnel will 

produce a drill program defining drill site location and depth so as to provide core representative 

of at least 70 percent of the total mined tonnage. The geological team developing the resource 

model will be required to establish the drill density and location necessary to achieve the 

representivity required for the PFS metallurgical test work program. It is probable that at least 30 

to 40 holes will need to be drilled which, in elevation and by depth, consider geology, boundaries, 

lithology, grade, mineralogy etc. With the 32 existing METCON drill holes, metallurgical results 

will be available covering 62 to 72 holes. 

Due to the size of this metallurgical program, it is expected that testing will contain elements of 

all of the following stages of metallurgical test work: 

• Stage I-Sample preparation 

• Stage II-Bottle roll and acid characterization testing 

• Stage III-Small column testing 

• Stage IV-Large column testing 

• Stage V-Final study preparation and recommendations for a Final Feasibility Study 

Note that this program may be revised prior to the PFS based on project needs and professional 

judgment. 

26.1.1 Stage I- Sample Preparation 

Once core is logged, the core will be composited into 50 foot interval composites versus depth in 

each hole. Each 50 foot composite will be sent to a metallurgical laboratory where the 

composites will be dried, screened, and the screen fractions retained separately. Each screen size 

will be assayed for total copper, oxide copper and cyanide sequential analysis, total iron and ICP. 

26.1.2 Stage II- Acid Bottle Roll and Acid Characterization Testing 

Each 50 foot composite will have a standard acid bottle roll test performed. With leaching at 100 

g/l sulfuric acid, variation in acid consumption is magnified. This variation can be evaluated 

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spatially and by depth to allow an increased understanding of global acid consumption in the 

deposit. 

Selected acid characterization testing will be performed. Static leach testing evaluates copper 

extraction and acid consumption versus crush sizes versus acid concentration over time. Crush 

sizes of one inch, 2 inches 4 inches and ROM will be leached at 3, 5, 10, 20 and 100 grams per 

liter for 40 days, maintaining the acid concentration within 90% of the designated acid 

concentration in each test. The PLS will be analyzed periodically over time to provide leach 

kinetics for copper extraction and acid consumption and the same for the gangue elements. 

Analytical testing of each acid characterization test provides an analysis of gangue dissolution 

occurring over specific time intervals by ICP analysis for selenium, uranium, aluminum, 

potassium, iron and troublesome anions of nitrates, chlorides and fluorides, all of which may be 

problematic and need to be addressed in SX/EW operations. 

26.1.3 Stage III- Small Column Leach Tests 

Selected 50 foot composites with head grades above the mine cut-off grade, but providing 

variable copper head grades, will be selected for small column testing by spatial location 

(horizontally and at depth). These tests will define metallurgical performance for copper 

extraction and acid consumption and variations spatially and by depth in the deposit 

26.1.4 Stage IV- Large Column Leach Tests 

Based on the data generated by the acid bottle roll, acid characterization testing, and the small 

column testing, master composites derived from the 50 foot individual composites will be 

prepared for large column testing to simulate ROM leaching. Some large columns will be run at 

a 20 foot height to evaluate the planned 20 foot leach pad lift in practice. 

26.1.5 Stage V- Study Preparation and Recommendations for a Final Feasibility 

Once the metallurgical study is completed, all data will be evaluated relative to copper extraction 

and acid consumption spatially in the deposit. Sufficient data will be available to resolve the 

questions raised in Section 26. Additional questions or observations generated from this work 

will be defined for evaluation during the feasibility study. 

From the metallurgical and block model data base generated by this study, extensive modeling of 

all elements of the project will be performed in an effort to select operational procedures in 

practice to optimize project performance. 

26.2 BUDGET AND SCHEDULE 

A budget and schedule has been developed for the follow on additional test work. The budget is 

in two phases, with the second phase dependent on positive results from the first phase. The first 

phase will consist of additional drilling to better define the resource, particularly in the area north 

of the MacArthur pit and at depth, and additional metallurgical drilling to obtain core samples for 

the metallurgical testing. This phase will also include an updated resource model and the nine 

month metallurgical test program as outlined above. 

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The second phase will consist of incorporating the additional resource information and 

metallurgical test results into a pre-feasibility study NI 43-101 Technical Report with an updated 

capital and operating cost estimate and an updated economic assessment of the project. This 

phase will develop additional preliminary engineering to establish the criteria for heap leach pad 

overall height, evaluation of the need for agglomeration, and market studies for the cost of sulfur 

delivered to the site. 

The budget and schedule for the two phase program is summarized in Table 26-1 below. 

Phase 1 

Phase 2 

Table 26-1: Budget for MacArthur Follow on Test Work 

Item Schedule Cost 

Additional Resource 

Drilling 100 holes 4 Months $1,100,000 

Additional Metallurgical 30 to 40 

Drilling 

Additional Metallurgical 

holes 4 Months $1,500,000 

Test Work 9 Months $1,000,000 

Subtotal Phase 1 9 Months $3,600,000 

Pre-feasibility Study & 

Technical Report 9 months $800,000 

Total Both Phases 18 months $4,400,000 

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27 REFERENCES 

Adams, D., 1987, MacArthur Au-Evaluation: Summary Report: unpublished report for Pangea 

Explorations, Inc.: 6p. 

Anaconda Collection-American Heritage Center, University of Wyoming, Laramie, Wyoming. 

Carneiro, R.R., MacArthur Project Preliminary Column Leach Study (Volume I). METCON 

Research. December 2011. 

Heatwole, D. A., 1972, Progress Report, Yerington Oxide Project – MacArthur Area, Lyon 

County, Nevada: unpublished private report for The Anaconda Company, 28p. 

Heatwole, D. A., 1978, Controls of Copper Oxide Mineralization, MacArthur Property, Lyon 

County, Nevada: Arizona Geological Society Digest, Volume XI, p 59-66. 

International Copper Study Group; Copper Marker Forecast 2011 – 2012; Press release dated 

October 4, 2011. 

Martin, D., 1989, MacArthur Metallurgical Review, Memo addressed to Mr. Bruce Riederer, 

Bateman Engineering 

Matson, E.J., 1952, Mac Arthur Copper Deposit, Lyon County, Nev., US Bureau of Mines 

Report of Investigation 4906, 47p. 

Moore, J. G., 1969, Geology and Mineral Deposits of Lyon, Douglas, and Ormsby Counties, 

Nevada: Nevada Bureau of Mines and Geology Bulletin 75, 45p. 

Nelson, P.H. and Van Voorhis, G.D., 1983, Estimation of sulfide content from induced 

polarization data, GEOPHYSICS, V.48, No. 1, pp. 62-75. 

Nevada Division of Wildlife (NDOW). 2012. Letter from timothy Herrick (Wildlife Resource 

Information) to Katie L. Dean (SRK): RE: MacArthur Exploration Project. March 16, 2012. 

Nevada Natural Heritage Program (NNHP). 2012. Letter from Eric S. Miskow (NNHP) to Katie 

L. Dean (SRK): RE: Data request received 08 March 2012. March 12, 2012. 

Pennington, J., Hartley, K., Lommen J., Willow, M., Scoping Study for the Re-mining and 

Processing of Residual Ore Stockpiles and Tailings, Yerington Copper Mine, Lyon County, 

Nevada. SRK Consulting (USA), Inc. March 14, 2012. 

Proffett, J.M. and Proffett, B.H., 1976, Stratigraphy of the Tertiary Ash-Flow Tuffs in the 

Yerington District, Nevada: Nevada Bureau of Mines and Geology, Report 27. 

Rozelle, J.W., MacArthur Copper Project, NI 43-101 Technical Report, Lyon County, Nevada, 

U.S.A. Tetra Tech MM Inc. January 21, 2011 

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Schmidt, R., 1996, Copper Mineralogy of Four Samples: Hazen Research, Inc.: unpublished 

private report for Arimetco, Inc., 10p. 

Tingley, J.V., Horton, R.C., and Lincoln, F.C., 1993, Outline of Nevada Mining History: Nevada 

Bureau of Mines and Geology, Special Publication 15, 48p. 

U.S. Bureau of Land Management (BLM). 2009. Environmental Assessment, Quaterra Alaska 

Inc. MacArthur Exploration Project. U.S. Department of the Interior, Bureau of Land 

Management, Carson City District, Sierra Front Field Office. October 2009. DOI-BLM-NV- 

C020-2010-0001-EA. 

USEPA, 2011, Supplemental Remedial Investigation Report, Arimetco Facilities Operable Unit 

8, Anaconda Copper Yerington Mine, Yerington, NV 

USEPA, 2008, Public Review Draft, Remedial Investigation Report, Arimetco Facilities 

Operable Unit 8, Anaconda Copper Yerington Mine, pages 170-172. 

USEPA, 2010, Data Summary Report for the Characterization of Vat Leach Tailings (VLT) 

Using X-Ray Fluorescence (XRF) - Yerington Mine Site 

USEPA, 2010, Historical Summary Report – Anaconda-Yerington Mine Site – Yerington, NV, 

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APPENDIX A: CERTIFICATE OF QUALIFIED PERSON (“QP”) AND CONSENT OF 

AUTHOR 

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APPENDIX A 

CERTIFICATE OF QUALIFIED PERSON (“QP”) AND 

CONSENT OF AUTHOR 


MAY 2012 

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CERTIFICATE OF QUALIFIED PERSON 

Rex Clair Bryan, Ph.D., MBA 

I, Rex Clair Bryan, Ph.D., MBA, do hereby certify that: 

1. I am currently employed by Tetra Tech MM, Inc. at: 

350 Indiana Street 

Suite 350 

Golden, Colorado 80401 

2. I graduated with a degree in Engineering (BS with honor) in 1971 and a MBA degree in 

1973 from the Michigan State University, East Lansing. In addition, I graduated from the 

Brown University with a degree in Geology in 1977, Providence, Rhode Island and The 

Colorado School of Mines, Golden, Colorado, with a graduate degree in Mineral 

Economics (Ph.D.) in 1980. 

3. I am a Registered Member (#411340) of the Society for Mining, Metallurgy, and 

Exploration, Inc. (SME). 

4. I have worked as a resource estimator and geostatistician for a total of thirty-one years 

since my graduation from university; as an employee of a leading geostatistical 

consulting company (Geostat Systems, Inc. USA), with large engineering companies 

such as Dames and Moore, URS, and Tetra Tech as a consultant for more than 30 years. 

5. I have read the definition of “qualified person” set out in National Instrument 43-101 

(“NI 43-101”) and certify that by reason of my education, affiliation with a professional 

association (as defined in NI 43-101) and past relevant work experience, I fulfill the 

requirements to be a “qualified person” for the purposes of NI 43-101. 

6. I am responsible for the preparation of the technical report (“Report”) titled MacArthur 

Copper Project, NI 43-101 Technical Report, Preliminary Economic Assessment, Lyon 

County, Nevada, USA dated May 23, 2012. I have visited the subject property on 

September 9-10, 2011 

7. I have either supervised the data collection, preparation, and analysis and/or personally 

completed an independent review and analysis of the data and written information 

contained in this Report. I am responsible for Sections 4-12 and 14 of this report. 

8. I have had no prior involvement with Quaterra Resources Inc. and/or the MacArthur 

Project and Property that is the subject of this Report. 

9. I am not aware of any material fact or material change with respect to the subject matter 

of the REPORT that is not reflected in the REPORT, the omission to disclose which 

makes the REPORT misleading. 

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10. I do not hold, nor do I expect to receive, any securities or any other interest in any 

corporate entity, private or public, with interests in the properties that are the subject of 

this report or in the properties themselves, nor do I have any business relationship with 

any such entity apart from a professional consulting relationship with the issuer, nor to 

the best of my knowledge do I have any interest in any securities of any corporate entity 

with property within a two (2) kilometer distance of any of the subject properties. 

11. I have read National Instrument 43-101 and Form 43-101F, and the REPORT has been 

prepared in compliance with that instrument and form. 

12. I consent to the filing of the Preliminary Economic Assessment NI 43-101 Technical 

Report with any stock exchanges or other regulatory authority and any publication by 

them, including electronic publication in the public company files on the websites 

accessible by the public, of the Preliminary Economic Assessment NI 43-101 Technical 

Report. 

Dated this 13 th Day of June, 2012 

________________________. 

Signature of Qualified Person 

“Rex Clair Bryan” . 

Print name of Qualified Person 

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CONSENT OF QUALIFIED PERSON 

I, Rex Clair Bryan, Ph.D., MBA, consent to public filing of the technical report entitled 

“MacArthur Copper Project, NI 43-101 Technical Report, Preliminary Economic Assessment, 

Lyon County, Nevada, USA” (Technical Report) dated May 23, 2012 by Quaterra Resources, 

Inc. 

I also consent to any extracts from or a summary of the Technical Report in the News Release 

dated May 23, 2012 by Quaterra Resources, Inc. 

I certify that I have read the News Release dated May 23, 2012 filed by Quaterra Resources, Inc. 

and that it fairly and accurately represents the information in the sections of the Technical Report 

for which I am responsible. 

Dated June 13, 2012 

Rex C. Bryan, Ph.D., MBA 

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APPENDIX B: PROPERTY LISTING 

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APPENDIX B 

PROPERTY LISTING 


MAY 2012 

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MacArthur Project Claims Listing 

BLM 

Lyon Co. 

Number Claim Name Reference Sec-Twp-Rng Location Date 

NMC932507 QT 1 388083 S14,15,22,23-T14N-R24E 5/24/2006 

NMC932508 QT 2 388084 S22,23-T14N-R24E 5/24/2006 

NMC932509 QT 3 388085 S14,23-T14N-R24E 5/24/2006 

NMC932510 QT 4 388086 S23-T14N-R24E 5/24/2006 

NMC932511 QT 5 388087 S14,23-T14N-R24E 5/24/2006 

NMC932512 QT 6 388088 S23-T14N-R24E 5/24/2006 

NMC932513 QT 7 388089 S14,23-T14N-R24E 5/24/2006 

NMC932514 QT 8 388090 S23-T14N-R24E 5/24/2006 

NMC932515 QT 9 388091 S14,23-T14N-R24E 5/24/2006 

NMC932516 QT 10 388092 S23-T14N-R24E 5/24/2006 

NMC932517 QT 11 388093 S14,23-T14N-R24E 5/24/2006 

NMC932518 QT 12 388094 S24-T14N-R24E 5/24/2006 

NMC932519 QT 13 388095 S14,23-T14N-R24E 5/24/2006 

NMC932520 QT 14 388096 S23-T14N-R24E 5/24/2006 

NMC932521 QT 15 388097 S14,23-T14N-R24E 5/24/2006 

NMC932522 QT 16 388098 S23-T14N-R24E 5/24/2006 

NMC932523 QT 17 388099 S14,23-T14N-R24E 5/24/2006 

NMC932524 QT 18 388100 S23-T14N-R24E 5/24/2006 

NMC932525 QT 19 388101 S13,14,23,24-T14N-R24E 5/24/2006 

NMC932526 QT 20 388102 S23,24-T14N-R24E 5/24/2006 

NMC932527 QT 21 388103 S13,24-T14N-R24E 5/23/2006 

NMC932528 QT 22 388104 S24-T14N-R24E 5/23/2006 

NMC932529 QT 23 388105 S13,24-T14N-R24E 5/23/2006 

NMC932530 QT 24 388106 S24-T14N-R24E 5/23/2006 

NMC932531 QT 25 388107 S13,24-T14N-R24E 5/23/2006 

NMC932532 QT 26 388108 S24-T14N-R24E 5/23/2006 

NMC932533 QT 27 388109 S13,24-T14N-R24E 5/23/2006 

NMC932534 QT 28 388110 S24-T14N-R24E 5/23/2006 

NMC932535 QT 29 388111 S13,24-T14N-R24E 5/23/2006 

NMC932536 QT 30 388112 S24-T14N-R24E 5/23/2006 

NMC932537 QT 31 388113 S13,24-T14N-R24E 5/23/2006 

NMC932538 QT 32 388114 S24-T14N-R24E 5/23/2006 

NMC932539 QT 33 388115 S13,24-T14N-R24E 5/23/2006 

NMC932540 QT 34 388116 S24-T14N-R24E 5/23/2006 

NMC932541 QT 35 388117 S13,24-T14N-R24E 5/23/2006 

NMC932542 QT 36 388118 S24-T14N-R24E 5/23/2006 

M3-PN110127 

23 May 2012 

Revision 0 242



NMC932543 QT 37 388119 

S13,24-T14N-R24E S18,19- 

T14N-R25E 

S24-T14N-R24E S19-T14N- 

5/23/2006 

NMC932544 QT 38 388120 

R25E 5/23/2006 

NMC932545 QT 39 388121 S18,19-T14N-R25E 5/23/2006 

NMC932546 QT 40 388122 S19-T14N-R25E 5/23/2006 

NMC932547 QT 41 388123 S18,19-T14N-R25E 5/23/2006 

NMC932548 QT 42 388124 S19-T14N-R25E 5/23/2006 

NMC932549 QT 43 388125 S18,19-T14N-R25E 5/23/2006 

NMC932550 QT 44 388126 S19-T14N-R25E 5/23/2006 

NMC932551 QT 45 388127 S18,19-T14N-R25E 5/23/2006 

NMC932552 QT 46 388128 S19-T14N-R25E 5/23/2006 

NMC932553 QT 47 388129 S18,19-T14N-R25E 5/23/2006 

NMC932554 QT 48 388130 S19-T14N-R25E 5/23/2006 

NMC932555 QT 49 388131 S18,19-T14N-R25E 5/23/2006 

NMC932556 QT 50 388132 S19-T14N-R25E 5/23/2006 

NMC932557 QT 51 388133 S18,19-T14N-R25E 5/25/2006 

NMC932558 QT 52 388134 S19-T14N-R25E 5/25/2006 

NMC932559 QT 53 388135 S17,18,19,20-T14N-R25E 5/25/2006 

NMC932560 QT 54 388136 S19,20-T14N-R25E 5/25/2006 

NMC932561 QT 55 388137 S22,23-T14N-R24E 5/24/2006 

NMC932562 QT 56 388138 S22,23,26,27-T14N-R24E 5/24/2006 

NMC932563 QT 57 388139 S23-T14N-R24E 5/24/2006 

NMC932564 QT 58 388140 S23,26-T14N-R24E 5/24/2006 

NMC932565 QT 59 388141 S23-T14N-R24E 5/24/2006 

NMC932566 QT 60 388142 S23,26-T14N-R24E 5/24/2006 

NMC932567 QT 61 388143 S23-T14N-R24E 5/24/2006 

NMC932568 QT 62 388144 S23,26-T14N-R24E 5/24/2006 

NMC932569 QT 63 388145 S23-T14N-R24E 5/24/2006 

NMC932570 QT 64 388146 S23,26-T14N-R24E 5/24/2006 

NMC932571 QT 65 388147 S23-T14N-R24E 5/24/2006 

NMC932572 QT 66 388148 S23,26-T14N-R24E 5/24/2006 

NMC932573 QT 67 388149 S23-T14N-R24E 5/24/2006 

NMC932574 QT 68 388150 S23,26-T14N-R24E 5/24/2006 

NMC932575 QT 69 388151 S23-T14N-R24E 5/26/2006 

NMC932576 QT 70 388152 S23,26-T14N-R24E 7/27/2006 

NMC932577 QT 71 388153 S23-T14N-R24E 5/26/2006 

NMC932578 QT 72 388154 S23,26-T14N-R24E 7/27/2006 

NMC932579 QT 73 388155 S23,24-T14N-R24E 5/26/2006 

NMC932580 QT 74 388156 S23,24,25,26-T14N-R24E 7/27/2006 

NMC932581 QT 75 388157 S24-T14N-R24E 5/26/2006 

M3-PN110127 

23 May 2012 

Revision 0 243



NMC932582 QT 76 388158 S24,25-T14N-R24E 7/27/2006 

NMC932583 QT 77 388159 S24-T14N-R24E 5/26/2006 

NMC932585 QT 79 388161 S24-T14N-R24E 5/23/2006 

NMC932587 QT 81 388163 S24-T14N-R24E 5/23/2006 

NMC932589 QT 83 388165 S24-T14N-R24E 5/23/2006 

NMC932591 QT 85 388167 S24-T14N-R24E 5/23/2006 

NMC932593 QT 87 388169 S24-T14N-R24E 5/23/2006 

NMC932595 QT 89 388171 S24-T14N-R24E 

S24-T14N-R24E S19-T14N- 

5/23/2006 

NMC932597 QT 91 388173 

R25E 5/23/2006 

NMC932599 QT 93 388175 S19-T14N-R25E 5/23/2006 

NMC932601 QT 95 388177 S19-T14N-R25E 5/23/2006 

NMC932603 QT 97 388179 S19-T14N-R25E 5/23/2006 

NMC932605 QT 99 388181 S19-T14N-R25E 5/23/2006 

NMC932607 QT 101 388183 S19-T14N-R25E 5/23/2006 

NMC932609 QT 103 388185 S19-T14N-R25E 5/25/2006 

NMC932610 QT 104 388186 S19,30-T14N-R25E 5/25/2006 

NMC932611 QT 105 388187 S19-T14N-R25E 5/25/2006 

NMC932612 QT 106 388188 S19,30-T14N-R25E 5/25/2006 

NMC932613 QT 107 388189 S19,20-T14N-R25E 5/25/2006 

NMC932614 QT 108 388190 S19,20,29,30-T14N-R25E 5/25/2006 

NMC932615 QT 109 388191 S20,29-T14N-R25E 5/25/2006 

NMC932616 QT 110 388192 S20,29-T14N-R25E 5/25/2006 

NMC932617 QT 111 388193 S26,27-T14N-R24E 5/26/2006 

NMC932618 QT 112 388194 S26,27-T14N-R24E 5/26/2006 

NMC932619 QT 113 388195 S26-T14N-R24E 5/26/2006 

NMC932620 QT 114 388196 S26-T14N-R24E 5/26/2006 

NMC932621 QT 115 388197 S26-T14N-R24E 5/26/2006 

NMC932622 QT 116 388198 S26-T14N-R24E 5/26/2006 

NMC932623 QT 117 388199 S26-T14N-R24E 5/26/2006 

NMC932639 QT 133 388215 S30-T14N-R25E 5/25/2006 

NMC932641 QT 135 388217 S29,30-T14N-R25E 5/25/2006 

NMC932642 QT 136 388218 S29,30-T14N-R25E 5/25/2006 

NMC932643 QT 137 388219 S29-T14N-R25E 5/25/2006 

NMC932644 QT 138 388220 S29-T14N-R25E 5/25/2006 

NMC932645 QT 139 388221 S29-T14N-R25E 5/25/2006 

NMC932646 QT 140 388222 S29-T14N-R25E 5/25/2006 

NMC932647 QT 141 388223 S26,27-T14N-R24E 5/26/2006 

NMC932648 QT 142 388224 S26,27-T14N-R24E 5/26/2006 

NMC932649 QT 143 388225 S26-T14N-R24E 5/26/2006 

M3-PN110127 

23 May 2012 

Revision 0 244



NMC932650 QT 144 388226 S26,35-T14N-R24E 5/26/2006 

NMC932651 QT 145 388227 S26-T14N-R24E 5/26/2006 

NMC932652 QT 146 388228 S26,35-T14N-R24E 5/26/2006 

NMC932658 QT 152 388234 S25,36-T14N-R24E 5/25/2006 

NMC932660 QT 154 388236 S25,36-T14N-R24E 5/25/2006 

NMC932662 QT 156 388238 S25,36-T14N-R24E 5/25/2006 

NMC932664 QT 158 388240 S25,36-T14N-R24E 

S25,36-T14N-R24E S30,31- 

5/25/2006 

NMC932666 QT 160 388242 

T14N-R25E 5/25/2006 

NMC932667 QT 161 388243 S30-T14N-R25E 5/25/2006 

NMC932668 QT 162 388244 S30,31-T14N-R25E 5/25/2006 

NMC932669 QT 163 388245 S30-T14N-R25E 5/25/2006 

NMC932670 QT 164 388246 S30,31-T14N-R25E 5/25/2006 

NMC932671 QT 165 388247 S30-T14N-R25E 5/25/2006 

NMC932672 QT 166 388248 S30,31-T14N-R25E 5/25/2006 

NMC932673 QT 167 388249 S30-T14N-R25E 5/25/2006 

NMC932674 QT 168 388250 S30,31-T14N-R25E 5/25/2006 

NMC932676 QT 170 388252 S30,31-T14N-R25E 5/25/2006 

NMC932677 QT 171 388253 S30-T14N-R25E 5/25/2006 

NMC932678 QT 173 388254 S29,30-T14N-R25E 5/25/2006 

NMC932679 QT 174 388255 S29,30-T14N-R25E 5/25/2006 

NMC932680 QT 175 388256 S29-T14N-R25E 5/25/2006 

NMC932681 QT 176 388257 S29-T14N-R25E 5/25/2006 

NMC932682 QT 177 388258 S34,35-T14N-R24E 5/25/2006 

NMC932683 QT 178 388259 S35-T14N-R24E 5/25/2006 

NMC932684 QT 179 388260 S35-T14N-R24E 5/25/2006 

NMC932685 QT 180 388261 S35-T14N-R24E 5/25/2006 

NMC932686 QT 181 388262 S35-T14N-R24E 5/25/2006 

NMC932687 QT 182 388263 S35-T14N-R24E 5/25/2006 

NMC932688 QT 183 388264 S35-T14N-R24E 5/25/2006 

NMC932689 QT 184 388265 S35-T14N-R24E 5/25/2006 

NMC932690 QT 185 388266 S35-T14N-R24E 5/25/2006 

NMC932691 QT 186 388267 S35-T14N-R24E 5/25/2006 

NMC932692 QT 187 388268 S35-T14N-R24E 5/25/2006 

NMC932693 QT 188 388269 S35-T14N-R24E 5/25/2006 

NMC932694 QT 189 388270 S35-T14N-R24E 5/25/2006 

NMC932695 QT 190 388271 S35-T14N-R24E 5/25/2006 

NMC932696 QT 191 388272 S35-T14N-R24E 5/25/2006 

NMC932697 QT 192 388273 S35-T14N-R24E 5/25/2006 

NMC932698 QT 193 388274 S35-T14N-R24E 5/25/2006 

M3-PN110127 

23 May 2012 

Revision 0 245



NMC932699 QT 194 388275 S35-T14N-R24E 5/25/2006 

NMC932700 QT 195 388276 S35-T14N-R24E 5/25/2006 

NMC932701 QT 196 388277 S35,36-T14N-R24E 5/25/2006 

NMC932702 QT 197 388278 S36-T14N-R24E 5/25/2006 

NMC932703 QT 198 388279 S36-T14N-R24E 5/25/2006 

NMC932704 QT 199 388280 S36-T14N-R24E 5/25/2006 

NMC932705 QT 200 388281 S36-T14N-R24E 5/25/2006 

NMC932706 QT 201 388282 S36-T14N-R24E 5/25/2006 

NMC932707 QT 202 388283 S36-T14N-R24E 5/25/2006 

NMC932708 QT 203 388284 S36-T14N-R24E 5/25/2006 

NMC932709 QT 204 388285 S36-T14N-R24E 5/25/2006 

NMC932710 QT 205 388286 S36-T14N-R24E 5/25/2006 

NMC932711 QT 206 388287 S36-T14N-R24E 5/25/2006 

NMC932712 QT 207 388288 S36-T14N-R24E 5/25/2006 

NMC932713 QT 208 388289 S36-T14N-R24E 5/25/2006 

NMC932714 QT 209 388290 S36-T14N-R24E 5/25/2006 

NMC932715 QT 210 388291 S36-T14N-R24E 

S36-T14N-R24E S31-T14N- 

5/25/2006 

NMC932716 QT 211 388292 

R25E 5/25/2006 

NMC932717 QT 212 388293 

S36-T14N-R24E S31-T14N- 

R25E 5/25/2006 

NMC932718 QT 213 388294 S31-T14N-R25E 5/25/2006 

NMC932719 QT 214 388295 S31-T14N-R25E 5/25/2006 

NMC932720 QT 215 388296 S31-T14N-R25E 5/25/2006 

NMC932721 QT 216 388297 S31-T14N-R25E 5/25/2006 

NMC932722 QT 217 388298 S31-T14N-R25E 5/25/2006 

NMC932723 QT 218 388299 S31-T14N-R25E 5/25/2006 

NMC932724 QT 219 388300 S31-T14N-R25E 5/25/2006 

NMC932725 QT 220 388301 S31-T14N-R25E 5/25/2006 

NMC932726 QT 221 388302 S31-T14N-R25E 5/25/2006 

NMC932727 QT 222 388303 S31-T14N-R25E 5/25/2006 

NMC932728 QT 223 388304 S31-T14N-R25E 5/25/2006 

NMC932729 QT 224 388305 S31-T14N-R25E 

S27-T14N-R24E S34-T14N- 

5/25/2006 

NMC983708 QT 251 423181 

R24E 1/30/2008 

NMC983709 QT 252 423182 

S27-T14N-R24E S34-T14N- 

R24E 1/30/2008 

NMC983710 QT 253 423183 S34-T14N-R24E 1/30/2008 

NMC983711 QT 254 423184 S34-T14N-R24E 1/30/2008 

NMC983712 QT 255 423185 S34-T14N-R24E 1/30/2008 

NMC983713 QT 256 423186 S34-T14N-R24E 1/30/2008 

M3-PN110127 

23 May 2012 

Revision 0 246



NMC983714 QT 257 423187 

S3-T13N-R24E S34-T14N- 

R24E 1/30/2008 

NMC983715 QT 258 423188 S3-T13N-R24E 

S3-T13N-R24E S34-T14N- 

1/30/2008 

NMC983716 QT 259 423189 

R24E 1/30/2008 

NMC983717 QT 260 423190 S3-T13N-R24E 

S2,3-T13N-R24E S34,35- 

1/30/2008 

NMC983718 QT 261 423191 

T14N-R24E 1/30/2008 

NMC983719 QT 262 423192 S2,3-T13N-R24E 

S2-T13N-R24E S35-T14N- 

1/30/2008 

NMC983720 QT 263 423193 

R24E 1/30/2008 

NMC983721 QT 264 423194 S2-T13N-R24E 

S2-T13N-R24E S35-T14N- 

1/30/2008 

NMC983722 QT 265 423195 

R24E 1/30/2008 

NMC983723 QT 266 423196 S2-T13N-R24E 

S2-T13N-R24E S35-T14N- 

1/30/2008 

NMC983724 QT 267 423197 

R24E 1/30/2008 

NMC983725 QT 268 423198 S2-T13N-R24E 

S2-T13N-R24E S35-T14N- 

1/30/2008 

NMC983726 QT 269 423199 

R24E 1/30/2008 

NMC983727 QT 270 423200 S2-T13N-R24E 

S2-T13N-R24E S35-T14N- 

1/30/2008 

NMC983728 QT 271 423201 

R24E 1/30/2008 

NMC983729 QT 272 423202 S2-T13N-R24E 

S2-T13N-R24E S35-T14N- 

1/30/2008 

NMC983730 QT 273 423203 

R24E 1/30/2008 

NMC983731 QT 274 423204 S2-T13N-R24E 

S2-T13N-R24E S35-T14N- 

1/30/2008 

NMC983732 QT 275 423205 

R24E 1/30/2008 

NMC983733 QT 276 423206 S2-T13N-R24E 1/30/2008 

NMC 963173 MP 1 412825 S26-T14N-R24E 8/9/2007 

NMC 963174 MP 2 412826 S26,35-T14N-R24E 8/9/2007 

NMC 963175 MP 3 412827 S26-T14N-R24E 8/9/2007 

NMC 963176 MP 4 412828 S26,35-T14N-R24E 8/9/2007 

NMC 963177 MP 5 412829 S26-T14N-R24E 8/9/2007 

NMC 963178 MP 6 412830 S26,35-T14N-R24E 8/9/2007 

NMC 963179 MP 7 412831 S26-T14N-R24E 8/9/2007 

NMC 963180 MP 8 412832 S26,35-T14N-R24E 8/9/2007 

NMC 963181 MP 9 412833 S26-T14N-R24E 8/9/2007 

NMC 963182 MP 10 412834 S26,35-T14N-R24E 8/9/2007 

NMC 963183 MP 11 412835 S26-T14N-R24E 8/9/2007 

NMC 963184 MP 12 412836 S26,35-T14N-R24E 8/9/2007 

M3-PN110127 

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NMC 963185 MP 13 412837 S25,26-T14N-R24E 8/9/2007 

NMC 963186 MP 14 412838 S25,26,35,36-T14N-R24E 8/9/2007 

NMC 963187 MP 15 412839 S25-T14N-R24E 8/9/2007 

NMC 963188 MP 16 412840 S25,36-T14N-R24E 8/9/2007 

NMC 963189 MP 17 412841 S25-T14N-R24E 8/9/2007 

NMC 963190 MP 18 412842 S25,36-T14N-R24E 8/9/2007 

NMC 963191 MP 19 412843 S25-T14N-R24E 8/9/2007 

NMC 963192 MP 20 412844 S25,36-T14N-R24E 8/9/2007 

NMC 963193 MP 21 412845 S25-T14N-R24E 8/9/2007 

NMC 963194 MP 22 412846 S25,36-T14N-R24E 8/9/2007 

NMC 963195 MP 23 412847 S25-T14N-R24E 8/9/2007 

NMC 963196 MP 24 412848 S25-T14N-R24E 8/9/2007 

NMC 963197 MP 25 412849 S25-T14N-R24E 8/9/2007 

NMC 963198 MP 26 412850 S25-T14N-R24E 8/9/2007 

MP 27 

S25-T14N-R24E S30-T14N- 

NMC 963199 

412851 

R25E 8/9/2007 

NMC 963200 MP 28 412852 S30-T14N-R25E 8/9/2007 

NMC 963201 MP 29 412853 S30-T14N-R25E 8/9/2007 

NMC 963202 MP 30 412854 S26-T14N-R24E 8/9/2007 

NMC 963203 MP 31 412855 S26-T14N-R24E 8/9/2007 

NMC 963204 MP 32 412856 S26-T14N-R24E 8/9/2007 

NMC 963205 MP 33 412857 S26-T14N-R24E 8/9/2007 

NMC 963206 MP 34 412858 S26-T14N-R24E 8/9/2007 

NMC 963207 MP 35 412859 S26-T14N-R24E 8/9/2007 

NMC 963208 MP 36 412860 S26-T14N-R24E 8/9/2007 

NMC 963209 MP 37 412861 S26-T14N-R24E 8/9/2007 

NMC 963210 MP 38 412862 S26-T14N-R24E 8/9/2007 

NMC 963211 MP 39 412863 S26-T14N-R24E 8/9/2007 

NMC 963212 MP 40 412864 S26-T14N-R24E 8/9/2007 

NMC 963213 MP 41 412865 S25,26-T14N-R24E 8/9/2007 

NMC 963214 MP 42 412866 S25,26-T14N-R24E 8/9/2007 

NMC 963215 MP 43 412867 S25-T14N-R24E 8/9/2007 

NMC 963216 MP 44 412868 S25-T14N-R24E 8/9/2007 

NMC 963217 MP 45 412869 S25-T14N-R24E 8/9/2007 

NMC 963218 MP 46 412870 S25-T14N-R24E 8/9/2007 

NMC 963219 MP 47 412871 S25-T14N-R24E 8/9/2007 

NMC 963220 MP 48 412872 S25-T14N-R24E 8/9/2007 

NMC 963221 MP 49 412873 S25-T14N-R24E 8/9/2007 

NMC 963222 MP 50 412874 S25-T14N-R24E 8/9/2007 

NMC 963223 MP 51 412875 S25-T14N-R24E 8/9/2007 

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NMC 963224 MP 52 412876 S25-T14N-R24E 8/9/2007 

NMC 963225 MP 53 412877 S25-T14N-R24E 8/9/2007 

NMC 963226 MP 54 412878 S25-T14N-R24E 8/9/2007 

NMC 963227 MP 55 412879 S25-T14N-R24E 8/9/2007 

NMC 963228 MP 56 412880 S25-T14N-R24E 8/9/2007 

NMC 963229 MP 57 412881 S25-T14N-R24E 8/9/2007 

NMC 963230 MP 58 412882 S25-T14N-R24E 8/9/2007 

MP 59 

S25-T14N-R24E S30-T14N- 

NMC 963231 

412883 

R25E 8/9/2007 

MP 60 

S25-T14N-R24E S30-T14N- 

NMC 963232 

412884 

R25E 8/9/2007 

NMC 963233 MP 61 412885 S30-T14N-R25E 8/9/2007 

NMC 963234 MP 62 412886 S30-T14N-R25E 8/9/2007 

NMC 963235 MP 63 412887 S30-T14N-R25E 8/9/2007 

NMC 963236 MP 64 412888 S30-T14N-R25E 8/9/2007 

NMC 963237 MP 65 412889 S30-T14N-R25E 8/9/2007 

NMC 963238 MP 66 412890 S30-T14N-R25E 8/9/2007 

NMC 963239 MP 67 412891 S30-T14N-R25E 8/9/2007 

NMC 963240 MP 68 412892 S30-T14N-R25E 8/9/2007 

NMC 963241 MP 69 412893 S30-T14N-R25E 8/9/2007 

NMC 963242 MP 70 412894 S30-T14N-R25E 8/9/2007 

NMC 963243 MP 71 412895 S30-T14N-R25E 8/9/2007 

NMC 963244 MP 72 412896 S30-T14N-R25E 8/9/2007 

NMC 963245 MP 73 412897 S24,25-T14N-R24E 8/9/2007 

NMC 963246 MP 74 412898 S24,25-T14N-R24E 8/9/2007 

NMC 963247 MP 75 412899 S24,25-T14N-R24E 8/9/2007 

NMC 963248 MP 76 412900 S24,25-T14N-R24E 8/9/2007 

NMC 963249 MP 77 412901 S24,25-T14N-R24E 8/9/2007 

NMC 963250 MP 78 412902 S24,25-T14N-R24E 8/9/2007 

NMC 963251 MP 79 412903 S24,25-T14N-R24E 8/9/2007 

MP 80 

S24,25-T14N-R24E S19,30- 

NMC 963252 

412904 

T14N-R25E 8/9/2007 

NMC 963253 MP 81 412905 S19,30-T14N-R25E 8/9/2007 

NMC 963254 MP 82 412906 S19,30-T14N-R25E 8/9/2007 

NMC 963255 MP 83 412907 S19,30-T14N-R25E 8/9/2007 

NMC 963256 MP 84 412908 S19,30-T14N-R25E 8/9/2007 

NMC 963257 MP 85 412909 S19,30-T14N-R25E 8/9/2007 

NMC1004075 AT 1 438843 S9,10,15,16-T14N-R24E 12/19/2008 

NMC1004076 AT 2 438844 S15,16-T14N-R24E 12/19/2008 

NMC1004077 AT 3 438845 S10,15-T14N-R24E 12/19/2008 

NMC1004078 AT 4 438846 S15-T14N-R24E 12/19/2008 

M3-PN110127 

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NMC1004079 AT 5 438847 S10,15-T14N-R24E 12/19/2008 

NMC1004080 AT 6 438848 S15-T14N-R24E 12/19/2008 

NMC1004081 AT 7 438849 S10,15-T14N-R24E 12/19/2008 

NMC1004082 AT 8 438850 S15-T14N-R24E 12/19/2008 

NMC1004083 AT 9 438851 S10,15-T14N-R24E 12/19/2008 

NMC1004084 AT 10 438852 S15-T14N-R24E 12/19/2008 

NMC1004085 AT 11 438853 S10,15-T14N-R24E 12/19/2008 

NMC1004086 AT 12 438854 S15-T14N-R24E 12/19/2008 

NMC1004087 AT 13 438855 S10,15-T14N-R24E 12/19/2008 

NMC1004088 AT 14 438856 S15-T14N-R24E 12/19/2008 

NMC1004089 AT 15 438857 S10,15-T14N-R24E 12/19/2008 

NMC1004090 AT 16 438858 S15-T14N-R24E 12/19/2008 

NMC1004091 AT 17 438859 S10,14,15-T14N-R24E 12/18/2008 

NMC1004092 AT 18 438860 S14,15-T14N-R24E 12/18/2008 

NMC1004093 AT 19 438861 S10,11,14-T14N-R24E 12/18/2008 

NMC1004094 AT 20 438862 S14-T14N-R24E 12/18/2008 

NMC1004095 AT 21 438863 S11,14-T14N-R24E 12/18/2008 

NMC1004096 AT 22 438864 S14-T14N-R24E 12/18/2008 

NMC1004097 AT 23 438865 S11,14-T14N-R24E 12/18/2008 

NMC1004098 AT 24 438866 S14-T14N-R24E 12/18/2008 

NMC1004099 AT 25 438867 S11,14-T14N-R24E 12/19/2008 

NMC1004100 AT 26 438868 S14-T14N-R24E 12/19/2008 

NMC1004101 AT 27 438869 S11,14-T14N-R24E 12/19/2008 

NMC1004102 AT 28 438870 S14-T14N-R24E 12/19/2008 

NMC1004103 AT 29 438871 S11,14-T14N-R24E 12/19/2008 

NMC1004104 AT 30 438872 S14-T14N-R24E 12/19/2008 

NMC1004105 AT 31 438873 S11,14-T14N-R24E 12/19/2008 

NMC1004106 AT 32 438874 S14-T14N-R24E 12/19/2008 

NMC1004107 AT 33 438875 S11,14-T14N-R24E 12/19/2008 

NMC1004108 AT 34 438876 S14-T14N-R24E 12/19/2008 

NMC1004109 AT 35 438877 S40131-T14N-R24E 12/19/2008 

NMC1004110 AT 36 438878 S13,14-T14N-R24E 12/19/2008 

NMC1004111 AT 37 438879 S12,13-T14N-R24E 12/19/2008 

NMC1004112 AT 38 438880 S13-T14N-R24E 12/19/2008 

NMC1004113 AT 39 438881 S12,13-T14N-R24E 12/19/2008 

NMC1004114 AT 40 438882 S13-T14N-R24E 12/19/2008 

NMC1004115 AT 41 438883 S12,13-T14N-R24E 12/19/2008 

NMC1004116 AT 42 438884 S13-T14N-R24E 12/19/2008 

NMC1004117 AT 43 438885 S12,13-T14N-R24E 12/19/2008 

NMC1004118 AT 44 438886 S13-T14N-R24E 12/19/2008 

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NMC1004119 AT 45 438887 S15,16-T14N-R24E 12/18/2008 

NMC1004120 AT 46 438888 S15,16,22-T14N-R24E 12/18/2008 

NMC1004121 AT 47 438889 S15-T14N-R24E 12/18/2008 

NMC1004122 AT 48 438890 S15,22-T14N-R24E 12/18/2008 

NMC1004123 AT 49 438891 S15-T14N-R24E 12/18/2008 

NMC1004124 AT 50 438892 S15,22-T14N-R24E 12/18/2008 

NMC1004125 AT 51 438893 S15-T14N-R24E 12/18/2008 

NMC1004126 AT 52 438894 S15,22-T14N-R24E 12/18/2008 

NMC1004127 AT 53 438895 S15-T14N-R24E 12/18/2008 

NMC1004128 AT 54 438896 S15,22-T14N-R24E 12/18/2008 

NMC1004129 AT 55 438897 S15-T14N-R24E 12/18/2008 

NMC1004130 AT 56 438898 S15,22-T14N-R24E 12/18/2008 

NMC1004131 AT 57 438899 S15-T14N-R24E 12/18/2008 

NMC1004132 AT 58 438900 S15-T14N-R24E 12/18/2008 

NMC1004133 AT 59 438901 S15-T14N-R24E 12/18/2008 

NMC1004134 AT 60 438902 S15-T14N-R24E 12/18/2008 

NMC1004135 AT 61 438903 S14,15-T14N-R24E 12/18/2008 

NMC1004136 AT 62 438904 S14,15-T14N-R24E 12/18/2008 

NMC1004137 AT 63 438905 S14-T14N-R24E 12/18/2008 

NMC1004138 AT 64 438906 S14-T14N-R24E 12/18/2008 

NMC1004139 AT 65 438907 S14-T14N-R24E 12/18/2008 

NMC1004140 AT 66 438908 S14-T14N-R24E 12/18/2008 

NMC1004141 AT 67 438909 S14-T14N-R24E 12/18/2008 

NMC1004142 AT 68 438910 S14-T14N-R24E 12/18/2008 

NMC1004143 AT 69 438911 S14-T14N-R24E 12/18/2008 

NMC1004144 AT 70 438912 S14-T14N-R24E 12/18/2008 

NMC1004145 AT 71 438913 S14-T14N-R24E 12/18/2008 

NMC1004146 AT 72 438914 S14-T14N-R24E 12/18/2008 

NMC1004147 AT 73 438915 S14-T14N-R24E 12/18/2008 

NMC1004148 AT 74 438916 S14-T14N-R24E 12/18/2008 

NMC1004149 AT 75 438917 S14-T14N-R24E 12/18/2008 

NMC1004150 AT 76 438918 S14-T14N-R24E 12/18/2008 

NMC1004151 AT 77 438919 S14-T14N-R24E 12/18/2008 

NMC1004152 AT 78 438920 S14-T14N-R24E 12/18/2008 

NMC1004153 AT 79 438921 S13,14-T14N-R24E 12/18/2008 

NMC1004154 AT 80 438922 S13,14-T14N-R24E 12/18/2008 

NMC1004155 AT 81 438923 S13-T14N-R24E 12/19/2008 

NMC1004156 AT 82 438924 S13-T14N-R24E 12/19/2008 

NMC1004157 AT 83 438925 S13-T14N-R24E 12/19/2008 

NMC1004158 AT 84 438926 S13-T14N-R24E 12/19/2008 

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NMC1004159 AT 85 438927 S13-T14N-R24E 12/19/2008 

NMC1004160 AT 86 438928 S13-T14N-R24E 12/19/2008 

NMC1004161 AT 87 438929 S13-T14N-R24E 12/19/2008 

NMC1004162 AT 88 438930 S13-T14N-R24E 12/19/2008 

NMC1004163 AT 89 438931 S13-T14N-R24E 12/19/2008 

NMC1004164 AT 90 438932 S13-T14N-R24E 12/19/2008 

NMC1004165 AT 91 438933 S13-T14N-R24E 12/19/2008 

NMC1004166 AT 92 438934 S13-T14N-R24E 12/19/2008 

NMC1004167 AT 93 438935 S13-T14N-R24E 12/19/2008 

NMC1004168 AT 94 438936 S13-T14N-R24E 12/19/2008 

NMC1004169 AT 95 438937 S13-T14N-R24E 12/18/2008 

NMC1004170 AT 96 438938 S13-T14N-R24E 12/18/2008 

NMC1004171 AT 97 438939 S13-T14N-R24E 12/18/2008 

NMC1004172 AT 98 438940 S13-T14N-R24E 12/18/2008 

NMC1004173 AT 99 438941 S22-T14N-R24E 12/18/2008 

NMC1004174 AT 100 438942 S22-T14N-R24E 12/18/2008 

NMC1004175 AT 101 438943 S22-T14N-R24E 12/18/2008 

NMC1004176 AT 102 438944 S22-T14N-R24E 12/18/2008 

NMC1004177 AT 103 438945 S22-T14N-R24E 12/18/2008 

NMC1004178 AT 104 438946 S22-T14N-R24E 12/18/2008 

NMC1004179 AT 105 438947 S22-T14N-R24E 12/18/2008 

NMC1004180 AT 106 438948 S22-T14N-R24E 12/18/2008 

NMC1004181 AT 107 438949 S15,22-T14N-R24E 12/18/2008 

NMC1004182 AT 108 438950 S22-T14N-R24E 12/18/2008 

NMC1004183 AT 109 438951 S15,22-T14N-R24E 12/18/2008 

NMC1004184 AT 110 438952 S22-T14N-R24E 12/18/2008 

NMC1004185 AT 111 438953 S15,22-T14N-R24E 12/18/2008 

NMC1004186 AT 112 438954 S22-T14N-R24E 12/18/2008 

NMC1004187 AT 113 438955 S15,22-T14N-R24E 12/18/2008 

NMC1004188 AT 114 

TAUBERT 

438956 S22-T14N-R24E 12/18/2008 

NMC891081 HILLS 343020 S24-T14N-R24E 2/15/2005 

NMC1054412 AT 115 483055 S9,10-T14N-R24E 9/9/2011 

NMC1054413 AT 116 483056 S9,10-T14N-R24E 9/9/2011 

NMC1054414 AT 117 483057 S10-T14N-R24E 7/29/2011 

NMC1054415 AT 118 483058 S10-T14N-R24E 7/29/2011 

NMC1054416 AT 119 483059 S10-T14N-R24E 7/29/2011 

NMC1054417 AT 120 483060 S10-T14N-R24E 7/29/2011 

NMC1054418 AT 121 483061 S10-T14N-R24E 7/29/2011 

NMC1054419 AT 122 483062 S10-T14N-R24E 7/29/2011 

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NMC1054420 AT 123 483063 S10-T14N-R24E 7/29/2011 

NMC1054421 AT 124 483064 S10-T14N-R24E 7/29/2011 

NMC1054422 AT 125 483065 S10-T14N-R24E 7/29/2011 

NMC1054423 AT 126 483066 S10-T14N-R24E 7/29/2011 

NMC1054424 AT 127 483067 S10-T14N-R24E 7/29/2011 

NMC1054425 AT 128 483068 S10-T14N-R24E 7/29/2011 

NMC1054426 AT 129 483069 S10-T14N-R24E 7/29/2011 

NMC1054427 AT 130 483070 S10-T14N-R24E 7/29/2011 

NMC1054428 AT 131 483071 S10-T14N-R24E 7/29/2011 

NMC1054429 AT 132 483072 S10-T14N-R24E 7/29/2011 

NMC1054430 AT 133 483073 S10,11-T14N-R24E 7/29/2011 

NMC1054431 AT 134 483074 S10,11-T14N-R24E 7/29/2011 

NMC1054432 AT 135 483075 S11-T14N-R24E 7/29/2011 

NMC1054433 AT 136 483076 S11-T14N-R24E 7/29/2011 

NMC1054434 AT 137 483077 S11-T14N-R24E 7/29/2011 

NMC1054435 AT 138 483078 S11-T14N-R24E 7/29/2011 

NMC1054436 AT 139 483079 S11-T14N-R24E 7/29/2011 

NMC1054437 AT 140 483080 S11-T14N-R24E 7/29/2011 

NMC1054438 AT 141 483081 S11-T14N-R24E 7/29/2011 

NMC1054439 AT 142 483082 S11-T14N-R24E 7/29/2011 

NMC1054440 AT 143 483083 S11-T14N-R24E 7/29/2011 

NMC1054441 AT 144 483084 S11-T14N-R24E 7/29/2011 

NMC1054442 AT 145 483085 S11-T14N-R24E 7/29/2011 

NMC1054443 AT 146 483086 S11-T14N-R24E 7/29/2011 

NMC1054444 AT 147 483087 S11-T14N-R24E 7/29/2011 

NMC1054445 AT 148 483088 S11-T14N-R24E 7/29/2011 

NMC1054446 AT 149 483089 S11,12-T14N-R24E 7/29/2011 

NMC1054447 AT 150 483090 S11,12-T14N-R24E 7/29/2011 

NMC1054448 AT 151 483091 S12-T14N-R24E 7/29/2011 

NMC1054449 AT 152 483092 S12-T14N-R24E 7/29/2011 

NMC1054450 AT 153 483093 S12-T14N-R24E 7/29/2011 

NMC1054451 AT 154 483094 S12-T14N-R24E 7/29/2011 

NMC1054452 AT 157 483095 S9,10-T14N-R24E 9/9/2011 

NMC1054453 AT 158 483096 S10-T14N-R24E 7/29/2011 

NMC1054454 AT 159 483097 S10-T14N-R24E 7/29/2011 

NMC1054455 AT 160 483098 S10-T14N-R24E 7/29/2011 

NMC1054456 AT 161 483099 S10-T14N-R24E 7/29/2011 

NMC1054457 AT 162 483100 S10-T14N-R24E 7/29/2011 

NMC1054458 AT 163 483101 S10-T14N-R24E 7/29/2011 

NMC1054459 AT 164 483102 S10-T14N-R24E 7/29/2011 

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NMC1054460 AT 165 483103 S10-T14N-R24E 7/29/2011 

NMC1054461 AT 166 483104 S10,11-T14N-R24E 7/29/2011 

NMC1054462 AT 167 483105 S2,11-T14N-R24E 7/29/2011 

NMC1054463 AT 168 483106 S2,11-T14N-R24E 7/29/2011 

NMC1054464 AT 169 483107 S2,11-T14N-R24E 7/29/2011 

NMC1054465 AT 170 483108 S2,11-T14N-R24E 7/29/2011 

NMC1054466 AT 171 483109 S2,11-T14N-R24E 7/29/2011 

NMC1054467 AT 172 483110 S2,11-T14N-R24E 7/29/2011 

NMC1054468 AT 173 483111 S2,11-T14N-R24E 7/29/2011 

NMC1054469 AT 174 483112 S2,11,12-T14N-R24E 7/29/2011 

NMC1054470 AT 175 483113 S1,2,11,12-T14N-R24E 7/29/2011 

NMC1054471 AT 176 483114 S1,12-T14N-R24E 7/29/2011 

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APPENDIX C: EXPLORATION HISTORY OF THE MACARTHUR OXIDE COPPER 

PROPERTY 

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APPENDIX C 

EXPLORATION HISTORY OF THE MACARTHUR 

OXIDE COPPER PROPERTY 

BY THE ANACONDA COMPANY, 1972 

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Exploration History of the MacArthur Oxide Copper Property 

by The Anaconda Company, 1972 

My name is David Heatwole; from 1971 to 1974 I was a Project Geologist for The Anaconda 

Company (Anaconda), stationed in Weed Heights, Nevada. My primary responsibility during 

this time period was the exploration of the MacArthur Oxide copper property. I personally: 

1. did the original geologic mapping of the property 

2. designed and supervised the execution of the trenching program 

3. mapped the trenches and supervised the sampling 

4. designed drill programs and supervised site locations 

5. supervised the drill program and logged cuttings 

6. posted geologic and assay data to maps and sections 

7. calculated first reserve estimates 

8. collected samples for metallurgical testing 

The following report documents my recollections of the implementation of exploration work 

done by Anaconda on the MacArthur property. I have supplemented my memory by the written 

reports referenced on the last page. 

SURVEYING 

Initial geologic work was done on enlarged USGS 15 minute topographic maps. To lay out the 

trenching program surveyors from the Yerington Mine established primary triangulation stations 

on the project. The stations were placed by triangulation with a transit from established USGS 

survey points and previous stations located by the mine. The triangulation stations allowed work 

at MacArthur to use the Yerington Mine Grid, a rectangular coordinate system based at the 

Yerington pit. 

Yerington Mine surveyors established elevation control on the property by transit using vertical 

angles from known elevation points. 

The MacArthur trenches were laid out on a N30E direction perpendicular to the geologic grain 

established in early mapping. The end lines of the trenches were located by transit and stadia 

rod. To guide the bulldozer, stakes were placed along surface trace of the trenches using tape 

and compass. 

Before the drilling began, the mine surveyors triangulated additional control points on the 

property. Drill sites were located by transit/stadia, and compass and tape from the triangulation 

stations. 

In 2007, I was able to locate a number of Anaconda drill holes in areas that had not been 

disturbed by Arimetco’s mining operations. 

TRENCHING 

A trenching program designed to systematically assay outcropping copper oxide mineralization 

was accomplished in the later half of 1971. The trenches were laid out on 200 foot intervals 

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using the survey methods outlined above. 10,500 feet of trenches were dug to a depth averaging 

5 feet. About 850 of these trenches were deepened to a depth of 15’ to demonstrate the affect of 

surface “super-leach” on oxide copper grades. 

The trenches were mapped geologically at scales of 1”= 20;1’ = 50 and 1’=100 ; the scale 

depending upon geologic complexity. Survey control for the geologic mapping was tape and 

compass tied to triangulation and stadia points. 

After geologic mapping, trenches were sampled on 10 intervals. Survey control for the sampling 

was the same as those previously established by the geologic mapping. Sample locations were 

recorded in numbered sample tag books giving each sample a unique sample number. 

The samples collected are best described as “irregular rock chip”. Anaconda field personnel 

using geology picks, supplemented by single jack and moil, chipped horizontal samples at chest 

height. Considerable care was taken to assure that all fine material was collected in the samples. 

A brief description of sample procedure: 

1. the sample face was cleaned using a dry brush 

2. a canvas tarp was placed at the foot of the trench wall 

3. the sample was cut taking care that all material fell on the tarp 

4. the sample was transferred from the tarp to a new canvas sample bag 

5. the unique sample tag was placed in bag and the bag was sealed using attached cloth 

ties 

6. the samples were delivered at the end of each day to the Yerington Mine assay lab. 

Assay results were usually available within 24 hours. Assay results were averaged by myself 

and posted by hand to a 1”= 100 plan map. At a later date the trench assays were digitized and 

became part of what is now known as the Metech MacArthur database. 

DRILLING 

In 1972 over 225 holes (33,000 feet vertical and 13,000 feet angle) were drilled on the prospect 

using open hole percussion and rotary methods. 82 percent of the drilling was done using a 

modified Gardner-Denver PR123J “Air-trac” percussion rig. Additional drilling was done in 

1973. 

The Air-trac rig was fitted with a sampling system designed by Anaconda’s Mining Research 

department for drilling friable ore minerals. The sampling system consisted of modified drill 

collar that allowed fine material to be routed to an industrial dust collector. Although the Airtrac 

drilling was done dry, nothing was discharged to the atmosphere; 100 percent of the material 

exiting the hole was collected. 

Samples were normally collected at 5 foot intervals. The coarse and fine fractions were 

combined on site and split using a Jones splitter. Samples were bagged and tagged on site by the 

drill crew. An Anaconda field person picked up the samples daily and transferred them to the 

Yerington Mine assay lab. A mining engineer from Anaconda’s Mining Research department 

was on site to supervise the Air-trac drilling for most of the program. Sample recovery was 

estimated by weighing samples on site and comparing the sample weight to a calculated 

theoretical weight based on the volume drilled. 

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Boyles Brothers Drilling company completed the remainder of the drilling (18 percent) using a 

standard dry rotary drill rig. Boyles also designed a special sample collector to capture fine 

discharge from the hole. The Boyles system was not as efficient as Anaconda’s, but was 

successful in collecting much of the fine material. Boyles’s samples were split, tagged and 

bagged on site and picked up daily by Anaconda personnel. 

A small number of samples from this drilling were sent to Chemical and Mineralogical services 

(CMS) in Salt Lake City. 

ASSAYING 

The majority of samples from the MacArthur Project were assayed in the Yerington Mine assay 

lab. The Yerington Mine lab specialized in copper assays providing assay services to the mine 

and mill. The Yerington Mine used the “short iodide method” for copper assays. Anaconda’s 

geology department routinely checked the Yerington Mine’s assays by submitting duplicate 

samples to CMS. 

Anaconda’s geological research laboratory in Tucson, Arizona did check assays using atomic 

absorption spectrophotometry on both the Yerington Mine and CMS. (See attached report by 

Vincent, 1972) 

Respectfully submitted, 

David Heatwole 

Yerington District Exploration Manager 

Quaterra Alaska Inc 

October 2008 

REFERENCES 

Heatwole, David, 1972. Progress Report and Drilling Proposal MacArthur Claims, Lyon County 

Nevada, January 1972, Anaconda Company unpublished report. 

Heatwole, David, 1972, Progress Report Yerington Oxide Project-MacArthur Area, Lyon 

County Nevada: December 1972, Anaconda Company unpublished report. 

Vincent, Harold, 1972, Assay Checks for Drill Hole Samples from MacArthur Prospect, 

October, 1972, unpublished Anaconda inter-office memo. 

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APPENDIX D: EXPLORATION DRILL HOLES WITH INTERCEPTS 

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APPENDIX D 

EXPLORATION DRILL HOLES WITH INTERCEPTS 


MAY 2012 

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QUATERRA RESOURCES INC. 

MacArthur Copper Project 

Drill Hole intercepts - through Dec 31, 2011 

Complete Intercept Table 

Angle Total From To Thickness 

Total 

Cu 

Drill Hole Brg / Dip Depth feet feet feet % 

* QMT-1 0º/-90º 300 0 145 145 0.22 

including 

50 120 70 0.31 

170 210 40 0.15 

QMT-1aR 0º/-90º 300 0 165 165 0.26 

including 

40 85 45 0.53 

180 200 20 0.19 

QMT-1bR 0º/-90º 300 0 135 135 0.33 

including 

20 125 105 0.38 

185 210 25 0.19 

300 350 50 0.47 

including 

300 335 35 0.55 

* QMT-2 210º/-55º 300 0 245 245 0.29 

including 

40 170 130 0.38 

QMT-2aR 210º/-55º 170 0 55 55 0.2 

75 165 90 0.26 

including 

95 165 70 0.29 

* QMT-3 0º/-90º 352.5 0 120 120 0.24 

220 275 55 0.19 

QMT-3aR 0º/-90º 400 0 120 120 0.17 

140 150 10 0.13 

* QMT-4 0º/-90º 422.3 37.7 84 46.3 0.5 

110.5 174 63.5 0.17 

186.7 228.7 42 0.22 

274 304.3 30.3 0.31 

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* QMT-5 195º/-57º 352 36.8 112 75.2 0.15 

182 206 24 0.21 

245 275 30 0.15 

QMT-5aR 210º/-55º 400 0 135 135 0.2 

230 255 25 0.1 

285 320 35 0.11 

370 400 30 0.21 

* QMT-6 0º/-90º 394.7 33 128 95 0.25 

173 188 15 0.18 

257 322 65 0.28 

including 

268 322 54 0.31 

* QMT-7 0º/-90º 424 0 24 24 0.28 

48.4 70.3 21.9 0.29 

74.2 116 41.8 0.92 

including 

77.3 93.2 15.9 1.77 

129.6 154 24.4 0.18 

184 224 40 0.14 

254 284 30 0.2 

334 356.5 22.5 0.3 

* QMT-8 0º/-90º 353 10 29 19 0.19 

49 84 35 0.19 

142.2 229 86.8 0.2 

258.3 316 57.7 0.15 

QMT-8aR 0º/-90º 400 0 20 20 0.51 

40 85 45 0.22 

150 170 20 0.52 

185 360 175 0.24 

including 

305 355 50 0.43 

* QMT-9 0º/-90º 244 9 81 72 0.34 

116.3 173 56.7 0.16 

193 244 51 0.14 

* QMT-10 0º/-90º 480 84 109 25 0.22 

129 334 205 0.42 

including 

including 

136 214 78 0.78 

274 330.2 56.2 0.27 

349 372 23 0.3 

389 399 10 0.27 

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QMT-10aR 30º/-55º 350 80 180 100 0.27 

including 

140 165 25 0.44 

300 350 50 0.13 

QMT-10bR 0º/-90º 350 80 145 65 0.78 

including 

85 135 50 0.91 

185 350 165 0.38 

including 

including 

190 230 40 0.63 

295 335 40 0.55 

* QMT-11 0º/-90º 284 0 120 120 0.18 

147 180 33 0.14 

214 284 70 0.24 

including 

229 284 55 0.27 

QMT-11aR 0º/-90º 300 15 135 120 0.19 

including 

65 105 40 0.25 

160 220 60 0.16 

240 300 60 0.21 

* QMT-12 0º/-90º 326 0 10 10 0.16 

55 189 134 0.21 

229 317 88 0.2 

QMT-12aR 0º/-90º 110 25 40 15 0.12 

60 110 50 0.18 

* QMT-13 0º/-90º 309.2 0 164 164 0.21 

180 216 36 0.25 

228.4 241.3 12.9 0.26 

277.4 290.5 13.1 0.24 

QMT-13aR 0º/-90º 300 0 240 240 0.27 

including 

30 185 155 0.3 

270 300 30 0.2 

* QMT-14 210º/-55º 360 5 123 118 0.31 

including 

36.2 80.5 44.3 0.55 

203 263 60 0.26 

including 

218 258 40 0.29 

303 338 35 0.29 

QMT14aR 0º/-90º 350 0 190 190 0.26 

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including 

including 

0 120 120 0.33 

215 235 20 0.13 

250 325 75 0.23 

250 290 40 0.33 

340 350 10 0.54 

QMT-14bR 210º/-55º 350 0 115 115 0.4 

including 

30 115 85 0.48 

155 175 20 0.19 

200 260 60 0.16 

290 350 60 0.17 

* QMT-15 0º/-90º 350 12.5 118 105.5 0.36 

including 

72 108 36 0.4 

183.3 288 104.7 0.19 

QMT-15aR 0º/-90º 350 15 115 100 0.21 

230 350 120 0.19 

including 

280 310 30 0.31 

* QMT-16 0º/-90º 455 36.5 199 162.5 0.18 

214 254 40 0.18 

277.9 339 61.1 0.14 

359 455 96 0.24 

including 

372.6 394 21.4 0.46 

QMT-16aR 0º/-90º 450 55 190 135 0.16 

230 265 35 0.16 

285 305 20 0.14 

355 450 95 0.23 

including 

370 405 35 0.35 

* QMT-17 0º/-90º 350 54 67.3 13.3 0.13 

87.3 208.9 121.6 0.16 

236 246 10 0.14 

QMT-17aR 0º/-90º 350 50 140 90 0.24 

including 

85 120 35 0.32 

170 180 10 0.12 

QMT-17bR 0º/-90º 350 60 80 20 0.19 

115 180 65 0.2 

240 250 10 0.13 

390 400 10 0.13 

M3-PN110127 

23 May 2012 

Revision 0 265



* QMT-18 0º/-90º 400 64 84 20 0.25 

112 189 77 0.2 

QMT-18aR 0º/-90º 350 80 90 10 0.16 

105 160 55 0.15 

190 200 10 0.15 

215 240 25 0.13 

310 325 15 0.13 

QMT-18bR 0º/-90º 350 115 175 60 0.18 

* QMT-19 0º/-90º 200 0 44 44 0.51 

including 

16 44 28 0.73 

* QME-1 0º/-90º 324 174 250 76 0.37 

including 

184 234 50 0.48 

* QME-2 0º/-90º 300.5 159 179 20 0.29 

258 300.5 42.5 0.27 

including 

263 288 25 0.4 

* QME-3 0º/-90º 303 63 166.5 103.5 0.16 

including 

72.5 93 20.5 0.28 

181.2 303 121.8 0.13 

* QME-4 0º/-90º 115 0 22.4 22.4 0.13 

QME-4aR 0º/-90º 230 0 20 20 0.13 

35 45 10 0.17 

70 115 45 0.14 

215 230 15 0.15 

* QME-5 210º/-50º 72.5 0 40 40 0.18 

QME-5aR 210º/-50º 80 0 80 80 0.23 

QME-6R 0º/-90º 200 40 50 10 0.11 

QME-8R 0º/-90º 340 0 10 10 0.13 

25 35 10 0.13 

70 100 30 0.18 

120 140 20 0.2 

195 265 70 0.17 

M3-PN110127 

23 May 2012 

Revision 0 266



QME-9R 0º/-90º 200 80 105 25 0.22 

130 140 10 0.13 

QME-10R 0º/-90º 400 0 20 20 0.44 

105 120 15 0.34 

QME-75R 0º/-90º 350 195 235 40 0.09 

QME-76R 0º/-90º 350 300 310 10 0.17 

QME-77R 0º/-90º 350 

No assays above cut-off 

QME-78R 0º/-90º 350 275 285 10 0.15 

315 330 15 0.14 

QME-79R 0º/-90º 350 210 250 40 0.31 

285 350 65 0.23 

QME-80R 0º/-90º 350 85 100 15 0.57 

185 245 60 0.34 

including 

QME-81R 0º/-90º 350 

190 205 15 0.79 


QMC-1aR 0º/-90º 340 85 190 105 0.16 

including 

160 190 30 0.27 

245 300 55 0.49 

QMC-1bR 270º/-45º 450 90 110 20 0.12 

185 255 70 0.13 

270 450 180 0.91 

including 

300 395 95 1.56 

QMC-4aR 0º/-90º 300 40 60 20 0.3 

QMC-4bR 270º/-45º 400 40 125 85 0.28 

160 275 115 0.24 

including 

including 

180 195 15 0.72 

305 400 95 0.57 

315 375 60 0.71 

QMC-21R 0º/-90º 400 165 205 40 0.26 

340 355 15 0.2 

M3-PN110127 

23 May 2012 

Revision 0 267



390 400 10 0.13 

QMC-22R 0º/-90º 400 0 40 40 0.44 

100 110 10 0.23 

345 355 10 0.35 

QMC-23R 0º/-90º 400 280 290 10 0.2 

340 365 25 1.25 

including 

340 355 15 1.97 

QMC-24R 0º/-90º 400 0 15 15 0.12 

40 105 65 0.17 

120 220 100 0.22 

QMC-25R 0º/-90º 350 70 80 10 0.1 

100 155 55 0.29 

including 

135 155 20 0.51 

305 330 25 0.12 

QMC-26R 0º/-90º 390 10 40 30 0.2 

65 95 30 0.29 

115 160 45 0.34 

including 

140 160 20 0.63 

200 220 20 0.14 

240 265 25 0.11 

QMC-26aR 180º/-45º 400 30 45 15 0.21 

75 95 20 0.24 

120 155 35 0.22 

175 205 30 0.25 

including 

185 195 10 0.48 

QMC-27R 0º/-90º 380 30 65 35 0.18 

80 170 90 0.13 

195 310 115 0.3 

including 

205 225 20 0.71 

* QMCC-1 0º/-90º 404 119 149 30 0.2 

179 204 25 0.14 

224 264 40 0.54 

289 303.6 14.6 0.36 

* QMCC-2 0º/-90º 454 34 115.3 81.3 0.21 

127 222.7 95.7 0.24 

M3-PN110127 

23 May 2012 

Revision 0 268



320 339 19 0.17 

351.2 416.8 65.6 0.18 

* QMCC-3 0º/-90º 400 107 334 227 0.22 

including 

286 334 48 0.41 

399.1 416.8 17.7 0.21 

* QMCC-4 0º/-90º 304 42.1 87 44.9 0.23 

including 

72 87 15 0.39 

* QMCC-5 0º/-90º 318.5 154 217.6 63.6 0.17 

* QMCC-6 0º/-90º 359 88.3 98.3 10 0.15 

* QMCC-7 0º/-90º 410 5 23 18 0.15 

89 134 45 0.19 

239 275.1 36.1 0.42 

* QMCC-8 0º/-90º 356 304 314 10 0.14 

* QMCC-9 0º/-90º 350 142.4 152.5 10.1 0.12 

254 264 10 0.14 

* QMCC-10 0º/-90º 325 95.5 144 48.5 0.44 

including 

119 144 25 0.74 

159 199 40 0.2 

* QMCC-11 0º/-90º 350 94 194 100 0.16 

including 

145 158.7 13.7 0.25 

* QMCC-12 0º/-90º 474 149 251.8 102.8 0.19 

281.7 333 51.3 0.14 

422.4 454.5 32.1 0.16 

* QMCC-13 0º/-90º 434 0 114 114 0.24 

including 

39.8 69 29.2 0.49 

* QMCC-14 0º/-90º 330 162.2 172.3 10.1 0.1 

241.5 251.7 10.2 0.15 

* QMCC-15 0º/-90º 375 182.8 286.7 103.9 0.16 

* QMCC-16 0º/-90º 325 5 78.2 73.2 0.14 

M3-PN110127 

23 May 2012 

Revision 0 269



including 

96.9 219.3 122.4 0.26 

143 156.4 13.4 0.84 

295 325 30 0.13 

* QMCC-17 0º/-90º 327.5 77.2 103 25.8 0.19 

277 290.5 13.5 0.12 

* QMCC-18 0º/-90º 369.5 77 97 20 0.13 

155.2 166.8 11.6 0.23 

182 212 30 0.22 

* QMCC-19 0º/-90º 360.4 274 287 13 0.13 

* QMCC-20 0º/-90º 333 163 183 20 0.15 

QM-001 0º/-90º 400 20 65 45 0.39 

including 

50 65 15 0.91 

150 160 10 0.16 

310 345 35 0.43 

QM-002 0º/-90º 400 0 15 15 0.19 

80 100 20 0.38 

including 

80 90 10 0.5 

270 330 60 0.12 

QM-003 0º/-90º 400 0 15 15 0.22 

40 175 135 0.38 

including 

75 140 65 0.52 

220 235 15 0.21 

260 290 30 0.24 

QM-004 0º/-90º 400 135 175 40 0.12 

QM-005 0º/-90º 450 210 220 10 0.11 

310 320 10 0.25 

QM-006 0º/-90º 400 110 155 45 0.28 

including 

130 150 20 0.5 

QM-007 0º/-90º 400 160 180 20 0.16 

220 280 60 0.43 

including 

220 250 30 0.75 

QM-008 0º/-90º 400 0 30 30 0.39 

M3-PN110127 

23 May 2012 

Revision 0 270



including 

0 20 20 0.47 

275 290 15 0.48 

QM-009 0º/-90º 400 40 95 55 0.18 

including 

45 60 15 0.28 

190 220 30 0.16 

QM-010 0º/-90º 870 25 60 35 0.17 

190 250 60 0.3 

370 385 15 0.42 

470 530 60 0.73 

including 

including 

QM-011 0º/-90º 355 

480 495 15 2.46 

575 625 50 0.4 

575 595 20 0.79 


QM-012 0º/-90º 400 145 155 10 0.12 

205 220 15 0.14 

QM-013 0º/-90º 290 15 50 35 0.24 

including 

40 50 10 0.59 

QM-014 0º/-90º 350 285 330 45 0.17 

including 

315 325 10 0.27 

QM-015 0º/-90º 400 160 230 70 0.28 

including 

190 210 20 0.55 

255 270 15 0.12 

QM-016 0º/-90º 390 65 80 15 0.17 

125 155 30 0.14 

175 230 55 0.23 

QM-017 0º/-90º 450 135 160 25 0.19 

175 230 55 0.3 

including 

200 225 25 0.49 

QM-018 30º/-45º 510 85 95 10 0.17 

140 200 60 0.15 

275 310 35 0.18 

355 420 65 0.32 

QM-019 210º/-60º 450 155 270 115 0.24 

M3-PN110127 

23 May 2012 

Revision 0 271



including 

180 260 80 0.27 

QM-020 0º/-45º 530 40 180 140 0.24 

215 340 125 0.22 

including 

240 300 60 0.32 

410 420 10 0.21 

QM-021 180º/-60º 450 85 110 25 0.15 

125 180 55 0.29 

including 

150 165 15 0.54 

315 420 105 0.22 

QM-022 0º/-90º 440 130 150 20 0.58 

including 

135 150 15 0.72 

295 305 10 0.15 

QM-023 210º/-60º 400 100 160 60 0.26 

including 

135 155 20 0.45 

290 300 10 0.14 

QM-024 210º/-70º 350 50 60 10 0.13 

115 125 10 0.13 

215 265 50 0.45 

QM-025 210º/-70º 520 100 180 80 0.21 

including 

160 170 10 0.41 

195 265 70 0.22 

including 

240 260 20 0.41 

* QM-026 0º/-90º 2,000.00 147 158.3 11.3 0.24 

860.5 880.5 20 0.35 

including 

865 875.5 10.5 0.56 

1,063.40 1,111.00 47.6 0.39 

QM-027 180º/-45º 540 0 30 30 0.1 

135 150 15 0.15 

210 240 30 0.36 

265 295 30 0.26 

310 335 25 0.17 

350 415 65 0.28 

430 540 110 0.17 

QM-028 0º/-60º 470 20 55 35 0.16 

165 205 40 0.13 

M3-PN110127 

23 May 2012 

Revision 0 272



including 

230 250 20 0.17 

270 365 95 0.14 

390 470 80 0.19 

430 445 15 0.31 

QM-029 180º/-45º 500 0 70 70 0.18 

230 270 40 0.24 

285 300 15 0.3 

375 465 90 0.32 

QM-030 180º/-45º 500 245 345 100 0.46 

360 475 115 0.38 

including 

360 385 25 0.62 

QM-031 0º/-60º 430 65 105 40 0.16 

225 275 50 0.2 

including 

240 250 10 0.42 

290 300 10 0.3 

QM-032 0º/-60º 500 150 230 80 0.11 

305 320 15 0.21 

405 460 55 0.15 

QM-033 270º/-45º 490 130 155 25 0.2 

175 415 240 0.33 

including 

including 

280 320 40 0.51 

405 415 10 1.53 

QM-034 90º/-45º 450 240 450 210 0.51 

including 

305 425 120 0.71 

QM-035 180º/-60º 800 15 90 75 0.16 

140 165 25 0.16 

270 290 20 0.25 

380 555 175 0.23 

including 

470 485 15 0.73 

* QM-036 0º/-90º 1,917.00 128 146 18 0.18 

198 255 57 0.31 

602.7 614 11.3 0.25 

QM-037 270º/-60º 900 15 70 55 0.18 

175 195 20 0.29 

480 525 45 0.23 

M3-PN110127 

23 May 2012 

Revision 0 273



QM-038 0º/-90º 800 0 190 190 0.2 

210 255 45 0.19 

340 385 45 0.39 

QM-039 180º/-45º 800 0 100 100 0.19 

including 

55 75 20 0.32 

265 275 10 0.54 

320 340 20 0.23 

365 395 30 0.17 

QM-040 270º/-60º 415 0 140 140 0.19 

including 

70 100 30 0.27 

190 260 70 0.23 

315 415 100 0.18 

including 

345 360 15 0.38 

* QM-041 0º/-90º 1,894.00 153 182.2 29.2 0.31 

233.5 284.5 51 0.51 

including 

271 284.5 13.5 1 

QM-042 0º/-45º 400 200 255 55 0.73 

including 

210 225 15 2.26 

280 325 45 0.29 

340 375 35 0.19 

QM-043 270º/-45º 620 250 265 15 0.16 

295 310 15 0.19 

345 390 45 0.31 

including 

365 375 10 0.71 

490 500 10 0.67 

QM-044 0º/-60º 965 0 60 60 0.22 

155 200 45 0.88 

including 

160 190 30 1.2 

225 255 30 0.41 

280 320 40 0.44 

435 460 25 0.3 

595 610 15 0.55 

820 840 20 0.3 

QM-045 150º/-45º 800 0 50 50 0.13 

175 250 75 0.14 

270 330 60 0.19 

M3-PN110127 

23 May 2012 

Revision 0 274



including 

including 

295 310 15 0.3 

360 375 15 0.24 

420 440 20 0.16 

575 600 25 0.27 

580 590 10 0.42 

750 780 30 0.1 

* QM-046 15º/-50º 1,502.00 228 253 25 0.23 

375 391 16 0.3 

791 805 14 0.19 

886 898.5 12.5 0.25 

983 993 10 0.4 

1,068.00 1,088.00 20 0.23 

1,279.00 1,355.00 76 0.74 

including 

1,283.00 1,300.00 17 2.27 

1,410.00 1,424.00 14 0.25 

1,468.00 1,478.00 10 0.29 

QM-047 0º/-90º 1,030.00 165 220 55 0.26 

245 290 45 0.33 

325 335 10 0.33 

365 375 10 0.28 

610 620 10 0.29 

635 680 45 0.32 

720 750 30 0.11 

770 785 15 0.22 

960 990 30 0.12 

QM-048 270º/-60º 1,000.00 70 90 20 0.23 

130 155 25 0.13 

170 185 15 0.17 

235 275 40 0.11 

525 540 15 0.35 

650 685 35 1.32 

including 

660 680 20 2.17 

720 750 30 0.3 

* QM-049 180º/-60º 1,478.00 264 294 30 0.61 

423.5 463 39.5 0.15 

732.2 747 14.8 0.28 

809 829 20 0.29 

QM-050 180º/-60º 800 40 75 35 0.21 

including 

50 60 10 0.43 

M3-PN110127 

23 May 2012 

Revision 0 275



115 145 30 0.27 

305 335 30 0.13 

QM-051 0º/-45º 400 280 290 10 0.13 

QM-052 180º/-45º 420 130 170 40 0.28 

including 

135 150 15 0.47 

185 200 15 0.26 

220 280 60 0.18 

including 

235 255 20 0.26 

QM-053 270º/-45º 490 50 60 10 0.15 

150 165 15 0.14 

QM-054 270º/-45º 480 190 205 15 0.25 

250 260 10 0.29 

295 345 50 0.59 

360 375 15 0.27 

QM-055 0º/-45º 500 0 115 115 0.17 

including 

70 85 15 0.31 

130 175 45 0.36 

including 

135 150 15 0.57 

195 230 35 0.2 

QM-056 270º/-45º 550 40 110 70 0.34 

155 225 70 0.16 

including 

200 215 15 0.34 

QM-057 0º/-90º 400 15 40 25 0.21 

80 175 95 0.3 

285 300 15 0.42 

QM-058 180º/-45º 450 0 30 30 0.22 

95 110 15 0.2 

125 265 140 0.41 

355 415 60 0.21 

QM-059 0º/-45º 450 70 80 10 0.17 

315 325 10 0.21 

QM-060 270º/-45º 400 50 85 35 0.15 

140 400 260 0.38 

including 

140 190 50 0.8 

M3-PN110127 

23 May 2012 

Revision 0 276



including 

140 160 20 1.48 

QM-061 0º/-90º 550 30 80 50 0.12 

155 250 95 0.19 

including 

190 225 35 0.26 

410 455 45 0.17 

QM-062 180º/-45º 500 0 10 10 0.19 

45 60 15 0.12 

QM-063 0º/-90º 500 0 30 30 0.13 

85 105 20 0.16 

215 240 25 0.17 

QM-064 0º/-45º 650 0 35 35 0.14 

340 370 30 0.22 

430 500 70 0.26 

including 

430 460 30 0.44 

QM-065 0º/-90º 520 295 310 15 0.31 

375 400 25 0.51 

485 505 20 0.17 

QM-066 0º/-90º 570 170 190 20 0.15 

395 545 150 0.26 

including 

440 450 10 1.2 

QM-067 0º/-90º 500 0 20 20 0.12 

110 230 120 0.25 

including 

175 225 50 0.42 

QM-068 0º/-90º 600 470 585 115 1.15 

including 

485 580 95 1.36 

QM-069 180º/-60º 450 80 90 10 0.2 

130 140 10 0.22 

QM-070 0º/-90º 490 315 325 10 0.22 

415 480 65 0.76 

including 

435 480 45 1.02 

QM-071 0º/-90º 470 60 125 65 0.25 

65 95 30 0.4 

M3-PN110127 

23 May 2012 

Revision 0 277



QM-072 0º/-90º 860 750 785 35 0.6 

including 

770 785 15 1.2 

QM-073 0º/-45º 520 95 110 15 0.15 

125 160 35 0.13 

270 335 65 0.17 

350 365 15 0.18 

QM-074 0º/-90º 460 20 100 80 0.16 

including 

85 100 15 0.3 

QM-075 0º/-90º 430 175 300 125 0.18 

including 

250 290 40 0.26 

355 395 40 0.19 

QM-076 0º/-45º 490 55 70 15 0.23 

430 460 30 0.33 

QM-077 0º/-90º 450 40 80 40 0.29 

145 190 45 0.25 

including 

150 165 15 0.43 

QM-078 180º/-45º 420 90 145 55 0.17 

195 255 60 0.45 

including 

210 230 20 0.83 

QM-079 180º/-45º 530 130 340 210 0.24 

including 

275 325 50 0.38 

including 

290 300 10 0.89 

QM-080 180º/-45º 500 0 230 230 0.23 

including 

30 100 70 0.33 

including 

195 220 25 0.32 

QM-081 180º/-45º 510 130 150 20 0.15 

440 460 20 0.26 

QM-082 0º/-90º 470 85 190 105 0.18 

210 355 145 0.14 

including 

240 260 20 0.25 

QM-083 0º/-90º 490 0 15 15 0.31 

100 170 70 0.15 

190 290 100 0.14 

M3-PN110127 

23 May 2012 

Revision 0 278



330 360 30 0.16 

375 415 40 0.17 

435 490 55 0.2 

QM-084 0º/-90º 450 0 85 85 0.24 

105 140 35 0.19 

180 200 20 0.19 

390 450 60 0.36 

QM-085 0º/-90º 490 0 95 95 0.43 

including 

0 60 60 0.59 

QM-086 0º/-90º 400 

QM-087 0º/-90º 500 



QM-088 180º/-60º 480 115 240 125 0.4 

including 

150 175 25 1.18 

255 335 80 0.34 

QM-089 0º/-90º 800 145 155 10 0.1 

190 210 20 0.23 

QM-090 0º/-90º 430 20 70 50 0.29 

90 240 150 0.24 

270 290 20 0.18 

QM-091 0º/-90º 600 70 110 40 0.21 

140 150 10 0.53 

230 240 10 0.36 

QM-092 0º/-90º 400 110 120 10 0.28 

235 250 15 0.16 

QM-093 0º/-90º 400 70 85 15 0.13 

QM-094 0º/-90º 500 20 30 10 0.15 

165 175 10 0.16 

310 335 25 0.32 

395 405 10 0.25 

430 440 10 0.45 

QM-095 0º/-90º 400 95 140 45 0.13 

M3-PN110127 

23 May 2012 

Revision 0 279



QM-096 0º/-90º 440 165 210 45 0.5 

230 270 40 0.26 

QM-097 0º/-90º 580 190 210 20 0.45 

370 380 10 0.24 

405 510 105 0.28 

QM-098 0º/-90º 570 140 150 10 0.27 

250 320 70 0.16 

350 385 35 0.26 

425 500 75 0.23 

* QM-099 0º/-90º 1529 


* QM-100 0º/-90º 1965 195 225 30 0.15 

1,203.50 1,268.50 65 0.58 

QM-101 0º/-90º 625 225 235 10 0.12 

425 435 10 0.56 

465 495 30 0.11 

545 575 30 0.27 

QM-102 0º/-90º 700 275 295 20 0.15 

405 415 10 0.22 

445 460 15 0.33 

QM-103 0º/-90º 450 185 225 40 0.32 

250 265 15 0.22 

QM-104 0º/-90º 500 285 295 10 0.165 

QM-105 0º/-90º 405 170 180 10 0.45 

225 305 80 0.33 

QM-106 0º/-90º 400 0 30 30 0.52 

80 95 15 0.2 

125 135 10 0.36 

185 205 20 0.77 

220 290 70 0.4 

QM-107 0º/-90º 425 110 125 15 0.13 

QM-108 0º/-90º 355 95 105 10 0.32 

M3-PN110127 

23 May 2012 

Revision 0 280



* QM-109 180º/-60º 1056 0 10 10 0.29 

350 360 10 0.11 

400 410 10 0.2 

QM-110 0º/-90º 400 170 195 25 0.17 

QM-111 0º/-90º 400 165 180 15 0.25 

195 235 40 0.16 

250 265 15 0.16 

QM-112 0º/-90º 400 15 165 150 0.15 

180 195 15 0.14 

315 350 35 0.35 

QM-113 0º/-90º 350 0 25 25 0.1 

60 200 140 0.21 

QM-114 0º/-90º 400 

QM-115 0º/-90º 600 



QM-116 0º/-90º 600 60 70 10 0.13 

330 340 10 0.1 

385 400 15 0.16 

QM-117 0º/-90º 400 


QM-118 0º/-90º 350 60 130 70 0.13 

165 175 10 0.14 

245 280 35 0.12 

QM-119 0º/-90º 550 30 80 50 0.15 

395 420 25 0.12 

440 450 10 0.16 

QM-120 0º/-90º 650 60 70 10 0.12 

85 130 45 0.12 

200 225 25 0.22 

285 385 100 0.47 

including 

295 335 40 0.91 

QM-121 0º/-90º 580 0 25 25 0.21 

M3-PN110127 

23 May 2012 

Revision 0 281



325 340 15 0.23 

380 410 30 0.11 

430 490 60 0.18 

520 555 35 0.25 

QM-122 0º/-90º 600 290 305 15 0.2 

QM-123 0º/-90º 670 100 110 10 0.15 

405 475 70 0.26 

505 520 15 0.4 

535 550 15 0.23 

570 580 10 0.63 

QM-124 0º/-90º 780 280 290 10 0.18 

360 390 30 0.14 

405 420 15 0.14 

QM-125 0º/-90º 500 70 80 10 0.14 

350 445 95 0.18 

475 500 25 0.78 

including 

485 500 15 1.21 

QM-126 0º/-60º 450 325 340 15 0.54 

QM-127 0º/-90º 700 380 390 10 0.28 

450 465 15 0.68 

500 510 10 0.14 

QM-128 0º/-90º 900 485 495 10 0.11 

570 580 10 0.21 

665 680 15 0.37 

835 850 15 0.18 

QM-129 0º/-90º 802.5 200 230 30 0.3 

250 300 50 0.23 

500 520 20 0.18 

730 745 15 0.39 

QM-130 180º/-60º 585 0 20 20 0.09 

QM-131 0º/-90º 965 


QM-132 260º/-60º 800 245 255 10 0.12 

QM-133 180º/-60º 550 195 270 75 0.36 

M3-PN110127 

23 May 2012 

Revision 0 282



285 405 120 0.27 

420 430 10 0.13 

QM-134 180º/-70º 455 25 40 15 0.14 

130 145 15 0.19 

170 185 15 0.12 

215 255 40 0.12 

QM-135 0º/-90º 475 125 160 35 0.12 

205 225 20 0.25 

255 270 15 0.19 

375 395 20 0.13 

QM-136 0º/-60º 600 90 120 30 0.39 

140 160 20 0.26 

280 295 15 0.17 

QM-137 60º/-70º 500 


QM-138 0º/-90º 550 130 140 10 0.2 

160 225 65 0.18 

including 

180 210 30 0.25 

QM-139 0º/-90º 400 90 140 50 0.18 

200 220 20 0.17 

255 285 30 0.15 

370 385 15 0.2 

QM-140 0º/-50º 500 60 135 75 0.18 

150 170 20 0.15 

230 245 15 0.9 

315 325 10 0.26 

QM-141 0º/-90º 530 0 60 60 0.33 

225 295 70 0.17 

315 330 15 0.16 

370 395 25 0.49 

QM-142 0º/-60º 400 0 10 10 0.12 

155 190 35 0.26 

320 355 35 0.49 

QM-143 0º/-90º 450 40 60 20 0.11 

120 280 160 0.17 

M3-PN110127 

23 May 2012 

Revision 0 283



325 350 25 0.12 

365 380 15 0.35 

QM-144 0º/-90º 500 115 225 110 0.29 

including 

175 220 45 0.49 

245 260 15 0.17 

QM-145 180º/-60º 675 0 50 50 0.28 

65 80 15 0.11 

105 135 30 0.18 

160 170 10 0.19 

QM-146 0º/-90º 500 25 35 10 0.17 

375 400 25 0.16 

QM-147 0º/-90º 400 20 45 25 0.21 

70 80 10 0.11 

100 130 30 0.13 

145 155 10 0.15 

360 400 40 0.18 

QM-148 0º/-90º 465 45 170 125 0.14 

200 250 50 0.11 

275 315 40 0.23 

335 345 10 0.18 

QM-149 180º/-60º 750 330 375 45 0.23 

480 495 15 0.21 

560 570 10 0.21 

600 610 10 0.68 

670 720 50 0.28 

QM-150 0º/-90º 600 45 80 35 0.17 

125 170 45 0.34 

310 335 25 0.32 

480 510 30 0.1 

555 575 20 0.43 

QM-151 0º/-90º 535 0 20 20 0.11 

35 55 20 0.13 

100 125 25 0.24 

155 400 245 0.26 

including 

245 305 60 0.38 

M3-PN110127 

23 May 2012 

Revision 0 284



QM-152 0º/-90º 500 200 210 10 0.24 

270 285 15 0.17 

340 350 10 0.14 

470 485 15 0.2 

QM-153 0º/-90º 515 100 225 125 0.16 

including 

100 125 25 0.26 

315 325 10 0.15 

450 495 45 0.16 

QM-154 0º/-90º 500 65 115 50 0.12 

165 245 80 0.25 

260 305 45 0.28 

325 335 10 0.25 

QM-155 180º/-60º 575 45 80 35 0.18 

455 465 10 0.15 

QM-156 0º/-90º 450 0 15 15 0.21 

35 45 10 0.5 

270 285 15 0.3 

340 350 10 0.14 

QM-157 0º/-90º 435 30 125 95 0.27 

including 

35 100 65 0.3 

140 150 10 0.18 

170 180 10 0.19 

330 385 55 0.15 

410 435 25 0.14 

QM-158 0 o /-70 490 10 45 35 0.27 

QM-159 0º/-60º 500 None 

QM-160 0º/-60º 500 0 30 30 0.39 

QM-161 180º/-75º 225 105 115 10 0.1 

QM-162 0º/-90º 990.0 360.0 440.0 80.0 0.24 

470.0 505.0 35.0 0.14 

* QM-163 0º/-90º 2,069.0 

No values above cut-off 

M3-PN110127 

23 May 2012 

Revision 0 285



* QM-164 0º/-90º 2,140.0 614.0 629.0 15.0 0.27 

685.5 782.0 96.5 0.34 

1,673.0 1,737.0 64.0 1.31 

including 

1,708.0 1,737.0 29.0 2.21 

* QM-165 0º/-90º 2,041.0 615.0 630.0 15.0 0.23 

860.0 985.0 125.0 0.28 

including 

915.0 930.0 15.0 0.94 

1,026.5 1,071.0 44.5 0.18 

1,089.0 1,110.0 21.0 0.15 

* QM-166 0º/-90º 2,685.5 1,268.0 1,280.0 12.0 0.31 

1,691.0 1,701.0 10.0 0.32 

1,776.5 1,786.8 10.3 0.35 

QM-167 0º/-90º 645.0 240.0 255.0 15.0 0.11 

285.0 300.0 15.0 0.13 

540.0 555.0 15.0 0.19 

QM-168 0º/-90º 500.0 245.0 275.0 30.0 0.13 

315.0 325.0 10.0 0.18 

345.0 360.0 15.0 0.15 

QM-169 0º/-90º 520.0 150.0 170.0 20.0 0.41 

240.0 410.0 170.0 0.18 

including 

370.0 410.0 40.0 0.25 

445.0 470.0 25.0 0.25 

QM-170 0º/-90º 700.0 290.0 315.0 25.0 0.11 

465.0 475.0 10.0 0.48 

QM-171 0º/-90º 730.0 100.0 120.0 20.0 0.24 

140.0 165.0 25.0 0.20 

285.0 370.0 85.0 0.17 

425.0 625.0 200.0 0.25 

including 

450.0 500.0 50.0 0.48 

QM-172 0º/-90º 540.0 0.0 35.0 35.0 0.22 

145.0 175.0 30.0 0.21 

including 

150.0 160.0 10.0 0.43 

QM-173 0º/-90º 500.0 70.0 155.0 85.0 0.18 

including 

90.0 130.0 40.0 0.23 

M3-PN110127 

23 May 2012 

Revision 0 286



QM-174 0º/-90º 600.0 


QM-175 180º/-45º 500.0 15.0 90.0 75.0 0.22 

195.0 205.0 10.0 0.24 

280.0 295.0 15.0 0.19 

425.0 435.0 10.0 0.46 

455.0 500.0 45.0 0.12 

QM-176 180º/-70º 980.0 50.0 60.0 10.0 0.11 

210.0 220.0 10.0 0.19 

235.0 255.0 20.0 0.18 

705.0 720.0 15.0 0.18 

* QM-177 180º/-60º 2,352.0 1,937.0 1,953.0 16.0 0.16 

2,015.0 2,025.0 10.0 0.16 

QM-178 180º/-45º 500.0 35.0 105.0 70.0 0.21 

150.0 225.0 75.0 0.38 

including 

150.0 185.0 35.0 0.58 

QM-179 180º/-45º 600.0 35.0 95.0 60.0 0.19 

120.0 155.0 35.0 0.13 

220.0 255.0 35.0 0.11 

QM-180 270º/-45º 535.0 0.0 65.0 65.0 0.10 

140.0 165.0 25.0 0.12 

190.0 220.0 30.0 0.89 

including 

190.0 210.0 20.0 1.17 

445.0 460.0 15.0 0.22 

QM-181 0º/-45º 500.0 0.0 45.0 45.0 0.15 

85.0 100.0 15.0 0.11 

160.0 265.0 105.0 0.25 

including 

200.0 230.0 30.0 0.34 

QM-182 180º/-60º 780.0 0.0 15.0 15.0 0.18 

235.0 265.0 30.0 0.21 

290.0 310.0 20.0 0.15 

335.0 380.0 45.0 0.13 

400.0 485.0 85.0 0.36 

QM-183 180º/-45º 500.0 60.0 80.0 20.0 0.12 

215.0 240.0 25.0 0.15 

M3-PN110127 

23 May 2012 

Revision 0 287



300.0 320.0 20.0 0.25 

360.0 400.0 40.0 1.37 

QM-184 90º/-45º 440.0 240.0 250.0 10.0 0.16 

345.0 425.0 80.0 0.17 

including 

345.0 360.0 15.0 0.31 

* QM-185 160º/-60º 473.5 154.0 169.0 15.0 0.13 

QM-186 90º/-45º 450.0 220.0 270.0 50.0 0.24 

QM-187 0º/-60º 440.0 180.0 200.0 20.0 0.35 

310.0 400.0 90.0 1.66 

including 

315.0 355.0 40.0 3.49 

QM-188 0º/-60º 440.0 75.0 120.0 45.0 0.19 

255.0 275.0 20.0 0.11 

QM-189 0º/-45º 520.0 130.0 190.0 60.0 0.14 

370.0 500.0 130.0 0.25 

including 

395.0 405.0 10.0 0.62 

QM-190 0º/-45º 700.0 215.0 275.0 60.0 0.17 

including 

225.0 235.0 10.0 0.37 

325.0 345.0 20.0 0.38 

535.0 650.0 115.0 0.48 

including 

535.0 580.0 45.0 0.71 

685.0 700.0 15.0 0.44 

QM-191 0º/-90º 400.00 195 275 80 0.33 

including 

215 270 55 0.41 

QM-192 90º/-45º 450.0 20.0 180.0 160.0 0.34 

195.0 260.0 65.0 0.26 

including 

225.0 240.0 15.0 0.58 

320.0 355.0 35.0 0.19 

QM-193 90º/-45º 500.0 60.0 210.0 150.0 0.62 

295.0 320.0 25.0 0.15 

420.0 490.0 70.0 0.17 

QM-194 90º/-45º 340.0 0.0 120.0 120.0 0.24 

M3-PN110127 

23 May 2012 

Revision 0 288



185.0 340.0 155.0 0.16 

QM-195 270º/-45º 400.0 0.0 15.0 15.0 0.20 

40.0 230.0 190.0 0.44 

including 

135.0 180.0 45.0 1.19 

QM-196 180º/-45º 300.0 0.0 125.0 125.0 0.20 

150.0 180.0 30.0 0.19 

QM-197 90º/-45º 300.0 40.0 90.0 50.0 0.18 

115.0 135.0 20.0 0.19 

QM-198 90º/-45º 400.0 0.0 45.0 45.0 0.11 

210.0 400.0 190.0 0.22 

including 

210.0 235.0 25.0 0.55 

QM-199 270º/-45º 600.0 45.0 60.0 15.0 0.11 

395.0 405.0 10.0 0.41 

QM-200 180º/-45º 480.0 0.0 15.0 15.0 0.39 

105.0 120.0 15.0 0.13 

280.0 315.0 35.0 0.31 

QM-201 270º/-45º 400.0 200.0 265.0 65.0 0.24 

QM-202 90º/-45º 500.0 0.0 45.0 45.0 0.30 

385.0 420.0 35.0 0.22 

QM-203 0º/-45º 400.0 0.0 40.0 40.0 0.22 

245.0 300.0 55.0 0.15 

QM-204 180º/-45º 600.0 170.0 190.0 20.0 0.12 

305.0 340.0 35.0 0.20 

390.0 470.0 80.0 0.88 

including 

400.0 435.0 35.0 1.73 

QM-205 180º/-60º 700.0 90.0 105.0 15.0 0.11 

205.0 215.0 10.0 0.22 

285.0 295.0 10.0 0.23 

375.0 490.0 115.0 0.33 

QM-206 180º/-50º 600.0 350.0 475.0 125.0 0.55 

540.0 565.0 25.0 0.21 

M3-PN110127 

23 May 2012 

Revision 0 289



QM-207 180º/-50º 600.0 170.0 180.0 10.0 1.19 

320.0 330.0 10.0 0.36 

535.0 565.0 30.0 0.13 

QM-208 270º/-60º 750.0 200.0 245.0 45.0 0.12 

265.0 295.0 30.0 0.16 

500.0 530.0 30.0 0.30 

670.0 750.0 80.0 0.60 

QM-209 180º/-60º 500.0 170.0 195.0 25.0 0.14 

330.0 350.0 20.0 0.16 

QM-210 90º/-45º 650.0 305.0 340.0 35.0 0.21 

425.0 460.0 35.0 0.33 

QM-211 180º/-60º 650.0 275.0 375.0 100.0 0.35 

QM-212 270º/-60º 700.0 365.0 455.0 90.0 0.47 

500.0 640.0 140.0 0.28 

QM-213 180º/-60º 700.0 215.0 260.0 45.0 0.15 

310.0 395.0 85.0 0.19 

540.0 555.0 15.0 0.47 

580.0 610.0 30.0 0.44 

including 

590.0 600.0 10.0 0.88 

QM-214 0º/-45º 400.0 220.0 235.0 15.0 0.16 

335.0 350.0 15.0 0.15 

QM-215 180º/-45º 500.0 170.0 185.0 15.0 0.13 

465.0 480.0 15.0 0.19 

QM-216 90º/-45º 460.0 265.0 415.0 150.0 0.23 

including 

265.0 315.0 50.0 0.38 

QM-217 0º/-90º 500.0 120.0 165.0 45.0 0.22 

210.0 275.0 65.0 0.14 

370.0 480.0 110.0 0.29 

including 

375.0 450.0 75.0 0.35 

QM-218 180º/-45º 400.0 250.0 265.0 15.0 0.41 

295.0 310.0 15.0 0.16 

QM-219 180º/-45º 520.0 245.0 255.0 10.0 0.45 

M3-PN110127 

23 May 2012 

Revision 0 290



including 

330.0 345.0 15.0 0.20 

435.0 520.0 85.0 0.20 

480.0 500.0 20.0 0.47 

QM-220 0º/-90º 500.0 140.0 160.0 20.0 0.18 

295.0 320.0 25.0 0.25 

375.0 405.0 30.0 0.25 

QM-221 180º/-50º 550.0 215.0 250.0 35.0 0.13 

285.0 310.0 25.0 0.13 

345.0 370.0 25.0 0.12 

425.0 475.0 50.0 0.25 

QM-222 180º/-45º 600.0 85.0 175.0 90.0 0.42 

including 

140.0 165.0 25.0 1.17 

QM-223 0º/-90º 550.0 160.0 170.0 10.0 0.24 

205.0 280.0 75.0 0.21 

300.0 400.0 100.0 0.53 

QM-224 180º/-45º 600.0 85.0 115.0 30.0 0.15 

155.0 200.0 45.0 0.21 

240.0 295.0 55.0 0.28 

400.0 440.0 40.0 0.14 

QM-225 180º/-55º 600.0 280.0 410.0 130.0 0.32 

445.0 465.0 20.0 0.21 

QM-226 180º/-45º 500.0 310.0 440.0 130.0 0.21 

including 

385.0 435.0 50.0 0.31 

QM-227 180º/-45º 500.0 325.0 335.0 10.0 0.16 

390.0 400.0 10.0 0.18 

QM-228 180º/-45º 500.0 210.0 320.0 110.0 0.18 

including 

235.0 260.0 25.0 0.35 

335.0 375.0 40.0 0.32 

410.0 430.0 20.0 0.17 

QM-229 0º/-90º 300.0 80.0 90.0 10.0 0.37 

230.0 280.0 50.0 0.55 

QM-230 90º/-45º 415.0 290.0 325.0 35.0 0.23 

M3-PN110127 

23 May 2012 

Revision 0 291



QM-231 90º/-45º 500.0 170.0 385.0 215.0 0.26 

including 

245.0 345.0 100.0 0.38 

415.0 435.0 20.0 0.16 

QM-232 0º/-90º 385.0 45.0 120.0 75.0 0.11 

QM-233 180º/-45º 500.0 60.0 75.0 15.0 0.14 

175.0 190.0 15.0 0.13 

225.0 270.0 45.0 0.16 

QM-234 180º/-60º 500.0 90.0 100.0 10.0 0.24 

365.0 385.0 20.0 0.14 

QM-235 0º/-90º 400.0 105.0 175.0 70.0 0.35 

including 

145.0 170.0 25.0 0.68 

QM-236 180º/-45º 550.0 190.0 205.0 15.0 0.14 

240.0 300.0 60.0 0.11 

315.0 385.0 70.0 0.37 

QM-237 0º/-90º 400.0 45.0 135.0 90.0 0.28 

including 

55.0 90.0 35.0 0.51 

195.0 215.0 20.0 0.13 

260.0 280.0 20.0 0.15 

305.0 315.0 10.0 0.22 

QM-238 180º/-50º 700.0 75.0 175.0 100.0 0.35 

including 

115.0 135.0 20.0 1.11 

195.0 295.0 100.0 0.21 

310.0 390.0 80.0 0.14 

405.0 435.0 30.0 0.29 

500.0 520.0 20.0 0.27 

QM-239 180º/-55º 500.0 215.0 225.0 10.0 0.19 

335.0 500.0 165.0 0.30 

QM-240 180º/-50º 550.0 160.0 175.0 15.0 0.15 

275.0 320.0 45.0 0.23 

375.0 525.0 150.0 0.53 

including 

420.0 465.0 45.0 1.01 

QM-241 180º/-55º 500.0 265.0 290.0 25.0 0.13 

305.0 425.0 120.0 0.52 

M3-PN110127 

23 May 2012 

Revision 0 292



QM-242 180º/-50º 500.0 55.0 120.0 65.0 0.11 

170.0 215.0 45.0 0.14 

255.0 380.0 125.0 0.35 

QM-243 180º/-45º 400.0 0.0 20.0 20.0 0.11 

35.0 160.0 125.0 0.23 

including 

50.0 80.0 30.0 0.41 

210.0 225.0 15.0 0.24 

QM-244 180º/-45º 670.0 0.0 35.0 35.0 0.18 

160.0 240.0 80.0 0.29 

including 

QM-245 180º/-45º 125.0 

185.0 235.0 50.0 0.39 


QM-246 180º/-45º 140.0 55.0 70.0 15.0 0.11 

QM-247 0º/-90º 360.0 45.0 60.0 15.0 0.24 

185.0 205.0 20.0 0.41 

QM-248 180º/-70º 400.0 35.0 75.0 40.0 0.25 

225.0 295.0 70.0 0.18 

345.0 380.0 35.0 0.12 

QM-249 180º/-60º 450.0 20.0 50.0 30.0 0.15 

125.0 170.0 45.0 0.43 

QM-250 0º/-90º 400.0 135.0 260.0 125.0 0.18 

275.0 305.0 30.0 0.18 

355.0 400.0 45.0 0.17 

QM-251 0º/-45º 500.0 280.0 310.0 30.0 0.14 

355.0 440.0 85.0 0.13 

QM-252 0º/-90º 400.0 70.0 80.0 10.0 0.14 

355.0 375.0 20.0 0.14 

QM-253 0º/-90º 400.0 250.0 300.0 50.0 0.12 

385.0 400.0 15.0 0.13 

QM-254 0º/-90º 400.0 45.0 80.0 35.0 0.19 

95.0 200.0 105.0 0.26 

215.0 240.0 25.0 0.20 

255.0 270.0 15.0 0.14 

M3-PN110127 

23 May 2012 

Revision 0 293



QM-255 0º/-90º 400.0 95.0 130.0 35.0 0.11 

195.0 270.0 75.0 0.20 

including 

255.0 265.0 10.0 0.64 

335.0 345.0 10.0 0.21 

QM-256 180º/-45º 500.0 210.0 220.0 10.0 0.19 

370.0 395.0 25.0 0.12 

QM-257 0º/-90º 500.0 220.0 250.0 30.0 0.18 

265.0 275.0 10.0 0.14 

QM-258 0º/-90º 450.0 0.0 60.0 60.0 0.17 

115.0 170.0 55.0 0.11 

420.0 450.0 30.0 0.11 

QM-259 180º/-45º 450.0 0.0 10.0 10.0 0.13 

170.0 180.0 10.0 0.12 

QM-260 90º/-45º 500.0 30.0 115.0 85.0 0.13 

255.0 270.0 15.0 0.23 

QM-261 0º/-45º 500.0 25.0 70.0 45.0 0.22 

255.0 300.0 45.0 0.19 

340.0 360.0 20.0 0.14 

385.0 420.0 35.0 0.17 

445.0 480.0 35.0 0.19 

QM-262 180º/-45º 400.0 0.0 20.0 20.0 0.22 

55.0 80.0 25.0 0.20 

105.0 120.0 15.0 0.26 

170.0 185.0 15.0 0.33 

205.0 220.0 15.0 0.21 

375.0 400.0 25.0 0.25 

QM-263 180º/-45º 450.0 0.0 35.0 35.0 0.14 

175.0 200.0 25.0 0.15 

330.0 380.0 50.0 0.26 

QM-264 180º/-55º 400.0 190.0 200.0 10.0 0.18 

230.0 285.0 55.0 0.19 

including 

235.0 245.0 10.0 0.39 

QM-265 180º/-55º 500.0 165.0 175.0 10.0 0.13 

M3-PN110127 

23 May 2012 

Revision 0 294



including 

240.0 310.0 70.0 0.19 

245.0 275.0 30.0 0.28 

QM-266 180º/-45º 450.0 70.0 200.0 130.0 0.20 

225.0 335.0 110.0 0.19 

400.0 410.0 10.0 0.25 

425.0 450.0 25.0 0.18 

QM-267 180º/-50º 450.0 10.0 30.0 20.0 0.26 

290.0 305.0 15.0 0.15 

320.0 345.0 25.0 0.12 

425.0 450.0 25.0 0.22 

QM-268 180º/-60º 300.0 75.0 85.0 10.0 0.25 

180.0 245.0 65.0 0.25 

QM-269 0º/-90º 350.0 0.0 20.0 20.0 0.42 

40.0 80.0 40.0 0.42 

including 

60.0 75.0 15.0 0.84 

110.0 120.0 10.0 0.46 

270.0 305.0 35.0 0.17 

320.0 330.0 10.0 0.20 

QM-270 180º/-45º 400.0 45.0 95.0 50.0 0.16 

160.0 170.0 10.0 0.24 

260.0 300.0 40.0 0.26 

340.0 360.0 20.0 0.30 

QM-271 90º/-45º 500.0 0.0 30.0 30.0 0.14 

55.0 125.0 70.0 0.16 

185.0 215.0 30.0 0.14 

265.0 315.0 50.0 0.21 

390.0 410.0 20.0 0.17 

QM-272 180º/-45º 300.0 225.0 275.0 50.0 0.14 

QM-273 180º/-60º 430.0 160.0 250.0 90.0 0.19 

including 

220.0 250.0 30.0 0.30 

275.0 340.0 65.0 0.28 

360.0 390.0 30.0 0.36 

405.0 430.0 25.0 0.15 

QM-274 0º/-90º 600.0 150.0 180.0 30.0 0.13 

225.0 255.0 30.0 0.12 

M3-PN110127 

23 May 2012 

Revision 0 295



395.0 410.0 15.0 0.19 

435.0 490.0 55.0 0.21 

535.0 560.0 25.0 0.38 

QM-275 180º/-45º 550.0 385.0 415.0 30.0 0.34 

440.0 500.0 60.0 0.46 

QM-276 0º/-90º 300.0 20.0 85.0 65.0 0.33 

120.0 145.0 25.0 0.11 

180.0 220.0 40.0 0.15 

QM-277 0º/-90º 250.0 0.0 100.0 100.0 0.19 

QM-278 0º/-90º 300.0 110.0 125.0 15.0 0.14 

155.0 245.0 90.0 0.19 

including 

190.0 220.0 30.0 0.31 

QM-279 0º/-90º 400.0 70.0 85.0 15.0 0.18 

150.0 165.0 15.0 0.17 

185.0 260.0 75.0 0.21 

QM-280 0º/-90º 400.0 125.0 220.0 95.0 0.16 

270.0 380.0 110.0 0.15 

including 

305.0 330.0 25.0 0.27 

QM-281 0º/-90º 480.0 85.0 175.0 90.0 0.15 

215.0 300.0 85.0 0.15 

440.0 460.0 20.0 0.19 

QM-282 0º/-45º 450.0 65.0 290.0 225.0 0.14 

305.0 365.0 60.0 0.17 

385.0 450.0 65.0 0.16 

including 

385.0 395.0 10.0 0.31 

QM-283 180º/-45º 595.0 95.0 120.0 25.0 0.14 

155.0 260.0 105.0 0.21 

385.0 395.0 10.0 0.25 

415.0 555.0 140.0 0.15 

QM-284 0º/-90º 450.0 0.0 150.0 150.0 0.29 

including 

70.0 105.0 35.0 0.60 

180.0 205.0 25.0 0.14 

295.0 450.0 155.0 0.20 

including 

355.0 390.0 35.0 0.31 

M3-PN110127 

23 May 2012 

Revision 0 296



QM-285 0º/-90º 400.0 0.0 75.0 75.0 0.30 

90.0 135.0 45.0 0.12 

150.0 175.0 25.0 0.18 

QM-286 0º/-90º 460.0 0.0 90.0 90.0 0.32 

115.0 140.0 25.0 0.22 

QM-287 0º/-90º 400.0 0.0 60.0 60.0 0.15 

120.0 235.0 115.0 0.20 

including 

150.0 165.0 15.0 0.35 

QM-288 0º/-90º 600.0 0.0 10.0 10.0 0.12 

55.0 100.0 45.0 0.26 

including 

QM-289 0º/-90º 400.0 

70.0 95.0 25.0 0.34 

145.0 185.0 40.0 0.11 

200.0 240.0 40.0 0.15 


QM-290 0º/-90º 400.0 270.0 305.0 35.0 0.14 

QM-291 0º/-90º 270.0 205.0 245.0 40.0 0.21 

QM-292 0º/-45º 500.0 20.0 55.0 35.0 0.16 

140.0 160.0 20.0 0.15 

180.0 285.0 105.0 0.19 

including 

235.0 275.0 40.0 0.32 

380.0 400.0 20.0 0.25 

QM-293 180º/-45º 500.0 115.0 215.0 100.0 0.21 

including 

115.0 135.0 20.0 0.56 

QM-294 90º/-45º 500.0 200.0 245.0 45.0 0.16 

365.0 400.0 35.0 0.19 

375.0 385.0 10.0 0.29 

QM-295 0º/-45º 400.0 70.0 145.0 75.0 0.15 

195.0 380.0 185.0 0.20 

including 

305.0 335.0 30.0 0.32 

QM-296 0º/-45º 370.0 110.0 125.0 15.0 0.12 

290.0 325.0 35.0 0.12 

M3-PN110127 

23 May 2012 

Revision 0 297



QM-297 0º/-90º 500.0 0.0 15.0 15.0 0.14 

155.0 175.0 20.0 0.45 

QM-298 0º/-90º 340.0 0.0 10.0 10.0 0.16 

70.0 100.0 30.0 0.15 

QM-299 0º/-90º 300.0 20.0 65.0 45.0 0.22 

105.0 145.0 40.0 0.15 

220.0 245.0 25.0 0.77 

QM-300 0º/-90º 490.0 80.0 105.0 25.0 0.20 

125.0 175.0 50.0 0.22 

QM-301 0º/-50º 700.0 360.0 375.0 15.0 0.16 

QM-302 0º/-90º 500.0 45.0 90.0 45.0 0.60 

320.0 340.0 20.0 0.16 

410.0 450.0 40.0 0.45 

QM-303 270º/-60º 500.0 175.0 195.0 20.0 0.20 

230.0 250.0 20.0 0.26 

QM-304 180º/-60º 800.0 195.0 210.0 15.0 0.15 

290.0 325.0 35.0 0.33 

395.0 405.0 10.0 0.21 

545.0 695.0 150.0 0.30 

QM-305 0º/-90º 500.0 100.0 170.0 70.0 0.18 

185.0 240.0 55.0 0.32 

465.0 480.0 15.0 0.30 

QM-306 0º/-90º 500.0 25.0 45.0 20.0 0.18 

65.0 135.0 70.0 0.17 

including 

including 

85.0 110.0 25.0 0.29 

205.0 255.0 50.0 0.23 

215.0 235.0 20.0 0.37 

QM-307 0º/-90º 300.0 0.0 15.0 15.0 0.23 

190.0 210.0 20.0 0.20 

QM-308 270º/-45º 250.0 30.0 40.0 10.0 0.12 

180.0 225.0 45.0 0.29 

QM-309 0º/-90º 300.0 135.0 145.0 10.0 0.24 

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230.0 250.0 20.0 0.17 

QM-310 0º/-90º 250.0 55.0 65.0 10.0 0.11 

80.0 90.0 10.0 0.12 

165.0 225.0 60.0 0.14 

QM-311 0º/-90º 300.0 55.0 65.0 10.0 0.19 

150.0 215.0 65.0 0.29 

including 

185.0 195.0 10.0 0.42 

260.0 280.0 20.0 0.51 

QM-312 0º/-90º 200.0 15.0 35.0 20.0 0.20 

QM-313 0º/-90º 400.0 165.0 180.0 15.0 0.14 

210.0 270.0 60.0 0.13 

335.0 400.0 65.0 0.17 

* Denotes core hole 

All intervals calculated using 0.1% copper cutoff 

REGULATORY NOTE: 

The samples from the MacArthur drilling program are prepared and assayed and by Skyline Assayers & Laboratories in 

Tucson, Arizona which is accredited by the American Association for Laboratory Accreditation (A2LA - certificate no. 

2953.01) and by ISO17025 compliant ALS Chemex Laboratories in Sparks, Nevada. 

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APPENDIX E: RESOURCE MODEL DRILL HOLE LISTING 

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APPENDIX E 

RESOURCE MODEL DRILL HOLE LISTING 


MAY 2012 

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Record # ⱡ DH Name Easting Northing 

Micromodel ® 

Deposit Modeling and Mine Planning System 

Version 7.00 

Elevation 

(ft) 

Bearing 

(degree) 

Plunge 

Depth 

(ft) 

Company-Core 

type* 

4 M-10-G-1 2437708.0 14687829.0 4776.4 0 90 250 Metech-RC 

5 M-10-G-2 2437708.0 14687829.0 4776.4 210 55 250 Metech-RC 

6 M-15-A1-1 2437079.5 14688117.0 4860.6 0 90 250 Metech-RC 

7 M-15-E1-1 2436744.5 14688338.0 4897.5 0 90 150 Metech-RC 

8 M-15-I-1 2437875.0 14687665.0 4754.4 0 90 200 Metech-RC 

9 M-15-L-1 2438123.5 14687525.0 4731.9 0 90 100 Metech-RC 

13 M-45-C50-1 2437264.2 14687687.0 4823.6 30 45 200 Metech-RC 

17 M-70-G-1 2437408.8 14687306.0 4759.5 0 90 250 Metech-RC 

18 M-70-G-2 2437408.8 14687306.0 4759.5 210 55 250 Metech-RC 

19 M-75-I-1 2437568.8 14687154.0 4777.5 0 90 150 Metech-RC 

21 M-80-G-1 2437363.8 14687246.0 4786.5 0 90 75 Metech-RC 

22 M-83-G-1 2437335.5 14687180.0 4805.5 0 90 60 Metech-RC 

23 M-83-G-2 2437335.5 14687180.0 4805.5 0 90 35 Metech-RC 

24 M-83-G-4 2437335.5 14687180.0 4805.5 210 55 250 Metech-RC 

26 M-90-G-4 2437319.2 14687170.0 4805.5 0 90 250 Metech-RC-Twin 

27 M0-A1-1 2437162.5 14688263.0 4859.4 0 90 195 Metech-RC 

28 M0-A1-3 2437155.8 14688266.0 4861.4 210 50 195 Metech-RC 

29 M0-B-1 2437341.2 14688164.0 4825.4 0 90 250 Metech-RC 

30 M0-C1-1 2437003.2 14688355.0 4856.4 0 90 250 Metech-RC 

31 M0-C50-1 2437487.2 14688072.0 4807.4 0 90 150 Metech-RC 

32 M0-C50-2 2437487.2 14688068.0 4807.4 210 55 75 Metech-RC 

33 M0-E-1 2437600.0 14688005.0 4794.4 0 90 150 Metech-RC 

34 M0-E1-1 2436824.2 14688464.0 4863.5 0 90 100 Metech-RC 

35 M0-E1-2 2436824.2 14688464.0 4863.5 30 55 100 Metech-RC 

36 M0-G1-1 2436648.5 14688566.0 4887.5 0 90 150 Metech-RC 

37 M0-K-1 2438123.8 14687695.0 4739.4 0 90 150 Metech-RC 

38 M105-A-1 2437724.2 14689155.0 4768.4 0 90 250 Metech-RC 

39 M105-A-2 2437724.2 14689155.0 4768.4 210 50 250 Metech-RC 

40 M105-B-1 2437886.8 14689053.0 4749.4 0 90 150 Metech-RC 

41 M105-C50-12 2438026.0 14688970.0 4743.4 0 90 450 Metech-RC 

42 M105-C50-3 2438022.8 14688970.0 4743.4 210 55 75 Metech-RC 

43 M105-E-1 2438132.0 14688910.0 4750.4 0 90 250 Metech-RC 

44 M105-E-2 2438116.0 14688896.0 4750.4 210 55 200 Metech-RC 

45 M105-I-1 2438487.0 14688686.0 4722.4 0 90 250 Metech-RC 

46 M105-I-2 2438474.0 14688689.0 4722.4 210 55 150 Metech-RC 

47 M105-K-1 2438659.2 14688601.0 4704.4 0 90 200 Metech-RC 

48 M110-G-1 2438336.8 14688844.0 4744.4 0 90 240 Metech-RC 

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49 M110-G-2 2438343.2 14688844.0 4746.4 210 55 250 Metech-RC 

50 M115-C-1 2438020.8 14689091.0 4738.7 0 90 160 Metech-RC 

51 M120-B-1 2437973.2 14689173.0 4740.4 0 90 200 Metech-RC 

52 M120-C50-1 2438089.5 14689102.0 4731.4 0 90 225 Metech-RC-Twin 


54 M120-E-1 2438208.8 14689036.0 4733.4 0 90 250 Metech-RC 

55 M120-I-1 2438580.0 14688806.0 4718.4 0 90 250 Metech-RC 

56 M120-I-2 2438580.0 14688809.0 4718.4 210 55 75 Metech-RC 

57 M120-K-1 2438739.0 14688727.0 4698.4 0 90 150 Metech-RC 

58 M120-Q-1 2439246.2 14688427.0 4674.4 0 90 125 Metech-RC 

59 M120-S-1 2439402.0 14688335.0 4643.4 0 90 350 Metech-RC 

60 M125-G-1 2438423.0 14688967.0 4733.4 0 90 425 Metech-RC 

61 M125-G-2 2438433.0 14688971.0 4733.4 30 50 240 Metech-RC 

62 M125-G-3 2438433.0 14688971.0 4733.4 210 55 250 Metech-RC 

63 M125-G-4 2438420.0 14688964.0 4733.4 0 90 125 Metech-RC 

64 M135-A-1 2437864.2 14689413.0 4748.4 0 90 150 Metech-RC 

65 M135-C50-1 2438165.8 14689232.0 4724.4 0 90 200 Metech-RC 

66 M135-E-1 2438288.5 14689158.0 4718.4 0 90 150 Metech-RC 

67 M135-E-2 2438285.2 14689158.0 4718.4 210 55 150 Metech-RC 

68 M135-I-1 2438643.5 14688935.0 4719.4 0 90 250 Metech-RC 

69 M135-K-1 2438812.2 14688853.0 4695.4 0 90 150 Metech-RC 

70 M135-K-2 2438809.0 14688849.0 4695.4 0 90 150 Metech-RC 

71 M140-G-1 2438496.2 14689096.0 4711.4 0 90 325 Metech-RC 

72 M140-G-2 2438496.0 14689110.0 4711.4 210 50 157 Metech-RC 

73 M140-G-3 2438492.8 14689109.0 4711.4 210 50 165 Metech-RC 

74 M148-E-1 2438365.2 14689275.0 4712.4 0 90 200 Metech-RC 

75 M15-A1-1 2437228.8 14688402.0 4860.6 0 90 300 Metech-RC 

76 M15-A1-2 2437228.8 14688402.0 4860.4 210 50 200 Metech-RC 

77 M15-A1-3 2437225.5 14688402.0 4860.4 300 45 150 Metech-RC 

78 M15-A1-4 2437225.5 14688402.0 4860.4 160 45 150 Metech-RC 

79 M15-A1-5 2437225.5 14688402.0 4860.4 250 45 150 Metech-RC 

80 M15-A1-6 2437225.5 14688402.0 4860.4 70 45 150 Metech-RC 

81 M15-B-1 2437414.5 14688290.0 4829.4 0 90 250 Metech-RC 

82 M15-B-2 2437414.5 14688290.0 4829.4 210 55 100 Metech-RC 

83 M15-C1-1 2437083.0 14688491.0 4838.4 0 90 250 Metech-RC 

84 M15-C1-2 2437079.5 14688494.0 4838.4 210 55 150 Metech-RC 

85 M15-C50-1 2437560.5 14688201.0 4800.4 0 90 100 Metech-RC 

86 M15-E-1 2437673.2 14688134.0 4785.4 0 90 150 Metech-RC 

87 M15-E1-1 2436887.5 14688599.0 4849.4 0 90 150 Metech-RC 

88 M15-G1-1 2436731.8 14688692.0 4879.4 0 90 150 Metech-RC 

89 M15-K-1 2438197.0 14687825.0 4724.4 0 90 200 Metech-RC 

90 M15-K-2 2438193.5 14687831.0 4724.4 210 45 150 Metech-RC 

91 M150-B-1 2438113.0 14689441.0 4727.4 0 90 200 Metech-RC 

92 M150-C50-1 2438245.8 14689361.0 4716.4 0 90 150 Metech-RC 

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93 M150-I-1 2438736.2 14689068.0 4697.4 0 90 250 Metech-RC 

94 M150-K-1 2438905.2 14688982.0 4690.4 0 90 175 Metech-RC 

95 M150-K-2 2438895.2 14688979.0 4690.4 210 55 100 Metech-RC 

96 M150-M-1 2439058.2 14688857.0 4672.4 0 90 150 Metech-RC 

97 M150-O-1 2439227.2 14688755.0 4662.4 0 90 200 Metech-RC 

98 M150-Q-1 2439386.0 14688689.0 4656.4 0 90 150 Metech-RC 

99 M150-S-1 2439565.0 14688587.0 4661.4 0 90 250 Metech-RC 

100 M155-G-1 2438566.0 14689235.0 4701.6 0 90 150 Metech-RC 

101 M157.5-I-1 2438744.5 14689160.0 4691.4 0 90 150 Metech-RC 

102 M165-A-1 2438013.8 14689672.0 4731.4 0 90 150 Metech-RC 

103 M165-E-1 2438428.2 14689427.0 4707.4 0 90 350 Metech-RC 

104 M165-E-2 2438441.2 14689427.0 4707.4 0 90 150 Metech-RC 

105 M165-I-1 2438779.5 14689219.0 4689.4 0 90 200 Metech-RC 

106 M165-K-1 2438985.0 14689102.0 4678.3 0 90 220 Metech-RC-Twin 

107 M165-K-2 2438994.8 14689105.0 4678.4 210 55 200 Metech-RC 

108 M165-K-3 2438955.2 14689118.0 4680.7 340 45 175 Metech-RC 

109 M165-M-1 2439134.8 14688986.0 4677.4 0 90 200 Metech-RC 

110 M165-O-1 2439300.5 14688884.0 4655.4 0 90 215 Metech-RC 

111 M165-Q-1 2439469.0 14688815.0 4651.4 0 90 200 Metech-RC 

112 M167.5-G 2438623.2 14689341.0 4693.4 0 90 200 Metech-RC 

113 M172.5-I-1 2438817.8 14689289.0 4685.5 0 90 150 Metech-RC-Twin 

114 M175-K-1 2439003.0 14689210.0 4673.8 0 90 200 Metech-RC 

115 M177-H-1 2438763.8 14689373.0 4688.3 0 90 200 Metech-RC 

116 M178-J-1 2438945.8 14689291.0 4679.8 0 90 175 Metech-RC 

117 M180-E-1 2438498.0 14689562.0 4711.4 0 90 200 Metech-RC 

118 M180-G-1 2438673.5 14689470.0 4700.4 0 90 200 Metech-RC 

119 M180-I-1 2438866.0 14689349.0 4681.4 0 90 235 Metech-RC 

120 M180-M-1 2439211.2 14689115.0 4668.4 0 90 150 Metech-RC 

121 M180-M-2 2439207.8 14689128.0 4666.4 160 45 150 Metech-RC 

122 M180-O-1 2439387.0 14689010.0 4654.4 0 90 150 Metech-RC 

123 M180-Q-1 2439545.5 14688944.0 4645.4 0 90 150 Metech-RC 

124 M180-S-1 2439720.8 14688875.0 4645.4 0 90 350 Metech-RC 

125 M180-S-2 2439370.0 14689039.0 4659.7 250 45 150 Metech-RC 

126 M183-L-1 2439147.2 14689213.0 4667.6 0 90 175 Metech-RC 

127 M185-K-1 2439054.0 14689293.0 4671.4 0 90 200 Metech-RC 

128 M187.5-I-1 2438897.5 14689415.0 4682.4 0 90 200 Metech-RC 

129 M190-G-1 2438738.0 14689540.0 4682.4 0 90 150 Metech-RC 

130 M195-I-1 2438935.8 14689481.0 4684.4 0 90 225 Metech-RC 

131 M195-K-1 2439107.8 14689399.0 4673.4 210 55 400 Metech-RC 

132 M195-K-2 2439107.8 14689412.0 4673.7 0 90 200 Metech-RC 

133 M195-L-1 2439191.0 14689332.0 4665.4 0 90 250 Metech-RC 

134 M195-M-1 2439287.8 14689248.0 4658.4 0 90 340 Metech-RC-Twin 

135 M195-M-2 2439281.0 14689248.0 4658.4 210 55 200 Metech-RC-Twin 

136 M195-M-3 2439277.2 14689277.0 4659.4 70 45 200 Metech-RC 

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137 M195-M-4 2439280.8 14689274.0 4659.4 160 45 150 Metech-RC 

138 M195-O-1 2439456.8 14689146.0 4650.4 0 90 150 Metech-RC 

139 M195-O-2 2439453.5 14689143.0 4650.4 210 55 250 Metech-RC 

140 M195-Q-1 2439625.5 14689070.0 4639.4 0 90 200 Metech-RC 

141 M20-G-1 2437877.5 14688082.0 4781.4 0 90 300 Metech-RC 

142 M20-G-2 2437877.5 14688082.0 4781.4 210 55 250 Metech-RC 

143 M200-W-1 2440168.0 14688811.0 4632.4 0 90 205 Metech-RC 

144 M202.5-I-1 2438974.2 14689541.0 4683.4 0 90 100 Metech-RC 

145 M204-G-1 2438798.2 14689659.0 4695.4 0 90 395 Metech-RC 

146 M205-G-2 2438814.5 14689666.0 4694.4 0 90 150 Metech-RC-Twin 

147 M205-J-1 2439082.8 14689523.0 4681.9 0 90 225 Metech-RC 

148 M205-L-1 2439252.2 14689398.0 4669.4 0 90 160 Metech-RC 

149 M210-I-1 2439012.2 14689611.0 4684.4 0 90 200 Metech-RC 


151 M210-M-1 2439350.5 14689410.0 4665.4 0 90 250 Metech-RC 

152 M210-M-2 2439347.2 14689407.0 4665.4 210 55 200 Metech-RC 

153 M210-M-3 2439344.0 14689410.0 4665.4 290 50 265 Metech-RC 

154 M210-O-1 2439499.8 14689321.0 4655.4 0 90 335 Metech-RC-Twin 

155 M210-O-2 2439493.0 14689320.0 4655.4 210 55 200 Metech-RC 

156 M210-Q-1 2439685.2 14689219.0 4640.2 0 90 200 Metech-RC 

157 M217-G-1 2438875.0 14689769.0 4684.4 0 90 200 Metech-RC 

158 M217-O-1 2439567.8 14689368.0 4652.4 0 90 175 Metech-RC 

159 M22.5-A1-1 2437277.0 14688458.0 4856.4 0 90 125 Metech-RC 

160 M225-I-1 2439076.0 14689717.0 4679.4 0 90 300 Metech-RC 

161 M225-K-1 2439261.2 14689635.0 4671.4 0 90 275 Metech-RC 

162 M225-M-1 2439420.5 14689532.0 4660.4 0 90 200 Metech-RC 

163 M225-M-2 2439420.5 14689532.0 4661.4 210 55 200 Metech-RC 

164 M225-M-3 2439427.0 14689529.0 4663.4 160 45 150 Metech-RC 

165 M225-O-1 2439599.5 14689434.0 4652.4 0 90 200 Metech-RC 

166 M225-O-2 2439596.0 14689434.0 4652.9 210 55 240 Metech-RC 

167 M225-Q-1 2439775.0 14689332.0 4640.4 0 90 250 Metech-RC 

168 M232-O-1 2439644.0 14689504.0 4652.4 210 55 200 Metech-RC 

169 M240-G-1 2438996.2 14689965.0 4687.4 0 90 200 Metech-RC 

170 M240-K-1 2439360.5 14689781.0 4662 0 90 355 Metech-RC 

171 M240-M-1 2439487.0 14689671.0 4657.4 0 90 340 Metech-RC 

172 M240-M-2 2439483.8 14689671.0 4657.4 210 55 150 Metech-RC 

173 M240-O-1 2439682.5 14689563.0 4651.4 0 90 200 Metech-RC 

174 M240-Q-1 2439851.5 14689468.0 4642.4 0 90 250 Metech-RC 

175 M240-S-1 2440007.2 14689375.0 4630.4 0 90 340 Metech-RC 

176 M255-K-1 2439414.2 14689897.0 4660.4 0 90 315 Metech-RC 

177 M255-M-1 2439583.2 14689791.0 4647.4 0 90 200 Metech-RC 

178 M255-O-1 2439759.0 14689693.0 4644.4 0 90 200 Metech-RC 

179 M255-Q-1 2439928.0 14689597.0 4636.4 0 90 200 Metech-RC 

180 M255-S-1 2440097.2 14689488.0 4624.4 0 90 200 Metech-RC 

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181 M270-O-1 2439822.2 14689828.0 4636.4 0 90 335 Metech-RC 

182 M270-Q-1 2440007.8 14689730.0 4634.4 0 90 200 Metech-RC-Twin 

183 M270-S-1 2440180.0 14689628.0 4622.4 0 90 250 Metech-RC-Twin 

184 M285-Q-1 2440074.5 14689852.0 4625.4 0 90 200 Metech-RC 

185 M285-S-1 2440250.0 14689750.0 4620.8 0 90 250 Metech-RC 

186 M30-A1-1 2437312.0 14688531.0 4849.8 0 90 340 Metech-RC 

187 M30-A1-2 2437312.0 14688531.0 4849.8 210 50 250 Metech-RC 

188 M30-A1-3 2437308.5 14688534.0 4849.4 160 45 150 Metech-RC 

189 M30-B-1 2437487.8 14688416.0 4808.4 0 90 250 Metech-RC 

190 M30-B-2 2437487.8 14688423.0 4808.4 210 55 200 Metech-RC 

191 M30-C1-1 2437153.0 14688614.0 4826.4 0 90 250 Metech-RC 

192 M30-C1-2 2437149.8 14688617.0 4826.4 210 55 100 Metech-RC 

193 M30-C50-1 2437630.5 14688330.0 4817.4 0 90 100 Metech-RC 

194 M30-C50-2 2437630.5 14688330.0 4817.4 210 55 50 Metech-RC 

195 M30-C50-3 2437623.8 14688340.0 4817.4 210 55 100 Metech-RC 

196 M30-E-1 2437756.2 14688257.0 4807.4 0 90 250 Metech-RC 

197 M30-E-2 2437756.2 14688257.0 4807.4 210 55 150 Metech-RC 

198 M30-E1-1 2436964.0 14688732.0 4841.4 0 90 150 Metech-RC 

199 M30-I-1 2438114.2 14688056.0 4747.4 0 90 200 Metech-RC 

200 M30-I-2 2438111.0 14688053.0 4747.4 210 45 150 Metech-RC 


202 M30-K-2 2438266.8 14687954.0 4730.4 210 55 150 Metech-RC 

203 M30-M-1 2438445.8 14687852.0 4717.4 0 90 200 Metech-RC 

204 M30-O-1 2438625.2 14687727.0 4679.6 0 90 250 Metech-RC 

205 M300-Q-1 2440151.0 14689975.0 4643.4 0 90 200 Metech-RC 

206 M35-B1-1 2437268.0 14688606.0 4830.4 0 90 175 Metech-RC 

207 M35-G-1 2437950.8 14688204.0 4770.4 0 90 300 Metech-RC 

208 M35-G-2 2437950.8 14688204.0 4770.4 210 55 245 Metech-RC 

209 M40-K-1 2438321.2 14688040.0 4731.4 210 55 250 Metech-RC 

210 M45-A-1 2437478.2 14688593.0 4818.6 0 90 150 Metech-RC 

211 M45-A1-1 2437381.8 14688664.0 4807.2 0 90 257 Metech-RC 

212 M45-A1-2 2437381.8 14688664.0 4807.4 210 45 250 Metech-RC 

213 M45-A1-3 2437381.8 14688664.0 4807.4 30 50 235 Metech-RC 

214 M45-A1-4 2437381.8 14688670.0 4807.4 70 45 175 Metech-RC 

215 M45-B-1 2437564.2 14688545.0 4807.4 0 90 250 Metech-RC 

216 M45-B-2 2437561.0 14688545.0 4807.4 210 55 75 Metech-RC 



219 M45-C1-3 2437219.5 14688746.0 4811.4 110 45 150 Metech-RC 

220 M45-C50-1 2437706.8 14688463.0 4797.4 0 90 250 Metech-RC 

221 M45-C50-2 2437703.5 14688462.0 4797.4 210 55 100 Metech-RC 

222 M45-E-1 2437826.2 14688389.0 4785.4 0 90 100 Metech-RC 

223 M45-E-2 2437826.2 14688389.0 4785.4 210 45 75 Metech-RC 

224 M45-E1-1 2437037.2 14688858.0 4848.4 0 90 200 Metech-RC 

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225 M45-I-1 2438177.5 14688189.0 4748.4 0 90 150 Metech-RC 

226 M45-K-1 2438353.2 14688080.0 4729.4 0 90 250 Metech-RC 

227 M5-G-1 2437791.2 14687952.0 4769.4 0 90 300 Metech-RC 

228 M5-G-2 2437791.2 14687952.0 4769.4 210 55 300 Metech-RC 

229 M50-D1-1 2437169.0 14688827.0 4820.4 0 90 145 Metech-RC 

230 M50-G-1 2438034.0 14688321.0 4776.4 0 90 300 Metech-RC 

231 M50-G-2 2438037.2 14688324.0 4775.4 210 55 250 Metech-RC 

232 M52.5-A1-1 2437420.2 14688713.0 4801.4 0 90 75 Metech-RC 

233 M60-A1-1 2437481.2 14688787.0 4789.4 0 90 450 Metech-RC 

234 M60-A1-2 2437481.2 14688787.0 4789.4 210 50 145 Metech-RC 

235 M60-A1-3 2437451.8 14688796.0 4789.4 30 50 165 Metech-RC 

236 M60-A1-4 2437451.8 14688796.0 4789.4 0 90 100 Metech-RC 

237 M60-B-1-2 2437644.0 14688675.0 4787.4 0 90 250 Metech-RC 

238 M60-B-3 2437644.0 14688675.0 4788.4 210 55 75 Metech-RC 

239 M60-C1-1 2437302.8 14688869.0 4815.4 0 90 150 Metech-RC 

240 M60-C1-2 2437289.8 14688865.0 4817.4 70 45 150 Metech-RC 

241 M60-C50-1 2437780.0 14688592.0 4789.4 0 90 100 Metech-RC 

242 M60-C50-2 2437780.0 14688592.0 4789.4 210 55 100 Metech-RC 

243 M60-E-1 2437902.8 14688518.0 4812.4 0 90 250 Metech-RC 

244 M60-E-2 2437899.2 14688522.0 4812.4 210 55 75 Metech-RC 

245 M60-E1-1 2437110.2 14688990.0 4847.4 0 90 150 Metech-RC 

246 M60-I-1 2438260.5 14688315.0 4772.8 0 90 250 Metech-RC 

247 M60-I-2 2438263.8 14688318.0 4772.4 210 55 75 Metech-RC 

248 M60-K-1 2438429.8 14688216.0 4731.4 0 90 100 Metech-RC 

249 M62-M-1 2438614.8 14688131.0 4706.4 0 90 40 Metech-RC 

250 M62-M-2 2438618.0 14688137.0 4706.4 0 90 400 Metech-RC 

251 M65-A-1 2437573.5 14688765.0 4793.6 0 90 150 Metech-RC 

252 M65-G-1 2438113.8 14688457.0 4796.4 0 90 500 Metech-RC 

253 M65-G-2 2438113.8 14688457.0 4796.3 0 90 200 Metech-RC 

254 M65-G-3 2438113.8 14688457.0 4796.4 210 55 250 Metech-RC 

255 M7.5-C1-1 2437041.5 14688418.0 4849.1 0 90 150 Metech-RC 

256 M7.5-D1-1 2436955.5 14688469.0 4843.4 0 90 150 Metech-RC 

257 M75-A-2 2437647.8 14688836.0 4776.1 0 90 145 Metech-RC 

258 M75-A1-1 2437548.0 14688916.0 4782.4 0 90 250 Metech-RC 

259 M75-A1-2 2437548.0 14688916.0 4782.4 210 55 255 Metech-RC 

260 M75-B-1 2437724.0 14688801.0 4771.4 0 90 250 Metech-RC 

261 M75-B-2 2437717.5 14688794.0 4771.4 210 55 150 Metech-RC 

262 M75-C1-1 2437386.0 14688992.0 4817.4 0 90 200 Metech-RC 

263 M75-C1-2 2437386.0 14688992.0 4817.4 210 45 150 Metech-RC 

264 M75-C50-1 2437860.0 14688715.0 4776.4 0 90 100 Metech-RC 

265 M75-C50-2 2437860.0 14688715.0 4776.4 210 55 100 Metech-RC 

266 M75-E-1 2437975.8 14688651.0 4808.4 0 90 250 Metech-RC 

267 M75-E-2 2437972.5 14688651.0 4808.4 210 55 75 Metech-RC 

268 M75-I-1 2438330.8 14688437.0 4767.4 0 90 275 Metech-RC-Twin 

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269 M75-I-2 2438327.2 14688437.0 4767.4 210 55 75 Metech-RC 

270 M75-K-1 2438516.0 14688345.0 4728.4 0 90 100 Metech-RC 

271 M80-G-1 2438180.8 14688569.0 4781.4 0 90 290 Metech-RC 

272 M80-G-2 2438180.8 14688569.0 4781.4 30 55 250 Metech-RC 

273 M80-G-3 2438180.8 14688569.0 4781.4 210 55 250 Metech-RC 

274 M90-A1-1 2437637.8 14689042.0 4782.4 0 90 210 Metech-RC 

275 M90-A1-2 2437637.8 14689042.0 4782.4 210 50 250 Metech-RC 

276 M90-B-1-2 2437807.0 14688927.0 4758.4 0 90 305 Metech-RC-Twin 

277 M90-B-3 2437807.0 14688930.0 4758.4 210 55 150 Metech-RC 

278 M90-C1-1 2437469.0 14689124.0 4819.4 0 90 200 Metech-RC 

279 M90-C50-1 2437936.5 14688844.0 4762.4 0 90 200 Metech-RC 

280 M90-C50-2 2437933.2 14688841.0 4762.4 210 55 100 Metech-RC 

281 M90-E-1 2438059.0 14688771.0 4775.4 0 90 295 Metech-RC 

282 M90-E-2 2438055.8 14688770.0 4775.4 210 55 70 Metech-RC 

283 M90-E-3 2438042.5 14688777.0 4775.4 210 55 150 Metech-RC 

284 M90-I-1 2438404.0 14688563.0 4738.4 0 90 300 Metech-RC 

285 M90-I-2 2438397.5 14688560.0 4738.4 210 55 200 Metech-RC 

286 M90-K-1 2438592.5 14688478.0 4712.4 0 90 100 Metech-RC 

287 M90-K-2 2438592.5 14688478.0 4712.4 210 55 75 Metech-RC 

288 M90-M-1 2438735.2 14688379.0 4697.4 0 90 200 Metech-RC 

289 M90-O-1 2438940.8 14688251.0 4692.4 0 90 400 Metech-RC-Twin 

290 M95-G-1 2438256.5 14688735.0 4763.4 0 90 300 Metech-RC-Twin 

291 M95-G-2 2438259.8 14688735.0 4763.4 210 55 250 Metech-RC 

321 QM-001 2435864.5 14691006.0 4944.4 0 90 400 Quaterra-RC-2009 





326 QM-006 2435344.0 14690076.0 4918 0 90 400 Quaterra-RC-2009 
















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346 QM-026 2438455.8 14692500.0 4921.2 156.9 89.2 2000 Quaterra-Core-2009 









355 QM-035 2433485.0 14687995.0 5129.2 186 62.74 800 Quaterra-RC-2009 


357 QM-037 2433474.0 14687992.0 5129.2 272.5 60.82 900 Quaterra-RC-2009 








365 QM-045 2433488.2 14688007.0 5129 119.2 45.63 800 Quaterra-RC-2009 

366 QM-046 2432612.5 14687995.0 5335.8 20 50 1502 Quaterra-Core-2009 




















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449 QM-129 2436871.0 14692033.0 4780 0 90 802.5 Quaterra-RC-2010 

























M3-PN110127 

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Revision 0 311







481 QM-161 2438336.2 14692038.0 4909.7 180 75 225 QM_2011 

482 QM-161a 2438336.2 14692038.0 4909.7 0 90 0 QM_2011 

483 QM-162 2438342.2 14691802.0 4946.5 0 90 990 QM_2011 

484 QM-163 2439915.8 14693459.0 4795.6 0 90 2069 QM_2011 

485 QM-164 2438105.5 14694458.0 4683.3 0 90 2140 QM_2011 

486 QM-165 2436834.0 14693581.0 4683 0 90 2041 QM_2011 

487 QM-166 2433894.2 14693602.0 4957.7 0 90 2685.5 QM_2011 

488 QM-167 2434369.2 14691096.0 4942.7 0 90 645 QM_2011 

489 QM-168 2433327.8 14689071.0 5162.8 0 90 500 QM_2011 

490 QM-169 2432844.2 14689104.0 5166.4 0 90 520 QM_2011 

491 QM-170 2431342.5 14688606.0 5271.1 0 90 700 QM_2011 

492 QM-171 2432267.0 14689065.0 5136.6 0 90 730 QM_2011 

493 QM-172 2432805.0 14686598.0 5188.2 0 90 540 QM_2011 

494 QM-173 2433819.0 14686530.0 5140.4 0 90 500 QM_2011 

495 QM-174 2433820.5 14686098.0 5073.7 0 90 600 QM_2011 

496 QM-175 2433803.8 14687582.0 5100 180 45 500 QM_2011 

497 QM-176 2435323.0 14688565.0 4964.7 180 70 980 QM_2011 

498 QM-177 2432122.2 14691213.0 4982.8 180 60 2352 QM_2011 

499 QM-178 2435875.0 14690993.0 4942.7 180 45 500 QM_2011 

500 QM-179 2436349.2 14691034.0 4944.3 180 45 600 QM_2011 

501 QM-180 2436835.2 14691070.0 4934.6 270 45 535 QM_2011 

502 QM-181 2436812.2 14691048.0 4934.9 0 45 500 QM_2011 

503 QM-182 2437772.8 14691565.0 4898.4 180 60 780 QM_2011 

504 QM-183 2438332.0 14690548.0 4761.5 180 45 500 QM_2011 

505 QM-184 2438825.0 14690869.0 4834.2 90 45 440 QM_2011 

506 QM-185 2433382.8 14690536.0 5064.6 160 60 473.5 QM_2011 

507 QM-186 2437418.0 14691009.0 4841.2 90 45 450 QM_2011 

508 QM-187 2439282.5 14691525.0 4772 0 60 440 QM_2011 

509 QM-188 2439823.8 14690545.0 4775 0 90 440 QM_2011 

510 QM-189 2439836.0 14691039.0 4784.3 0 45 520 QM_2011 

511 QM-190 2438847.2 14691584.0 4839.6 0 45 700 QM_2011 

512 QM-191 2438834.5 14690551.0 4774.7 0 90 400 QM_2011 

513 QM-192 2439509.5 14689325.0 4640 90 45 450 QM_2011 

514 QM-193 2439210.5 14689559.0 4679 90 45 500 QM_2011 

515 QM-194 2439324.8 14689048.0 4613.9 90 45 340 QM_2011 

516 QM-195 2438203.0 14688556.0 4707.4 270 45 400 QM_2011 

517 QM-196 2437324.5 14688561.0 4782 180 45 300 QM_2011 

518 QM-197 2437324.2 14688562.0 4781.7 90 45 300 QM_2011 

519 QM-198 2435845.0 14691573.0 4846.7 90 45 400 QM_2011 

520 QM-199 2432330.2 14688595.0 5211.9 270 45 600 QM_2011 

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521 QM-200 2433826.2 14687099.0 5122.4 180 45 480 QM_2011 

522 QM-201 2436303.0 14688067.0 4848.4 270 45 400 QM_2011 

523 QM-202 2433361.2 14686674.0 5221.3 90 45 500 QM_2011 

524 QM-203 2433320.2 14687090.0 5172.5 0 45 400 QM_2011 

525 QM-204 2439271.5 14691481.0 4779.1 180 45 600 QM_2011 

526 QM-205 2439324.2 14692068.0 4856.2 180 60 700 QM_2011 

527 QM-206 2439090.8 14692046.0 4843.4 180 50 600 QM_2011 

528 QM-207 2438581.2 14692020.0 4848.1 180 50 600 QM_2011 

529 QM-208 2438336.5 14692027.0 4921.4 270 60 750 QM_2011 

530 QM-209 2438580.5 14691541.0 4923.2 180 60 500 QM_2011 

531 QM-210 2438297.8 14691572.0 4976.7 90 45 650 QM_2011 

532 QM-211 2438307.8 14691562.0 4976.8 180 60 650 QM_2011 

533 QM-212 2438298.8 14691586.0 4976.6 270 60 700 QM_2011 

534 QM-213 2438091.2 14691554.0 4956.7 180 60 700 QM_2011 

535 QM-214 2438809.8 14690061.0 4707.6 0 45 400 QM_2011 

536 QM-215 2438285.2 14690013.0 4794.3 180 45 500 QM_2011 

537 QM-216 2438283.5 14690004.0 4795 90 45 460 QM_2011 

538 QM-217 2438339.2 14689321.0 4721.6 0 90 500 QM_2011 

539 QM-218 2438586.2 14690554.0 4766.8 180 45 400 QM_2011 

540 QM-219 2440078.0 14691013.0 4784.3 180 45 520 QM_2011 

541 QM-220 2439585.5 14691014.0 4831.2 0 90 500 QM_2011 

542 QM-221 2439580.0 14691022.0 4831.1 180 50 550 QM_2011 

543 QM-222 2438847.8 14691548.0 4827.1 180 45 600 QM_2011 

544 QM-223 2438836.0 14691800.0 4824.4 0 90 550 QM_2011 

545 QM-224 2439086.0 14691547.0 4814.9 180 45 600 QM_2011 

546 QM-225 2439596.8 14692061.0 4816 180 55 600 QM_2011 

547 QM-226 2439604.5 14691549.0 4719.5 180 45 500 QM_2011 

548 QM-227 2440325.0 14691027.0 4740.6 180 45 500 QM_2011 

549 QM-228 2439570.0 14690561.0 4770.1 180 45 500 QM_2011 

550 QM-229 2440325.0 14690548.0 4694.9 0 90 300 QM_2011 

551 QM-230 2438365.0 14690551.0 4753.8 90 45 415 QM_2011 

552 QM-231 2438818.8 14690547.0 4774.7 90 45 500 QM_2011 

553 QM-232 2438080.8 14690551.0 4775.5 0 90 385 QM_2011 

554 QM-233 2438076.0 14690547.0 4792.1 180 45 500 QM_2011 

555 QM-234 2440071.5 14690049.0 4668.6 180 60 500 QM_2011 

556 QM-235 2439066.0 14690075.0 4700.1 0 90 400 QM_2011 

557 QM-236 2439085.5 14690564.0 4762.9 180 45 550 QM_2011 

558 QM-237 2439867.8 14689550.0 4635.3 0 90 400 QM_2011 

559 QM-238 2439318.0 14690991.0 4862.6 180 50 700 QM_2011 

560 QM-239 2439177.8 14691048.0 4862.2 180 55 500 QM_2011 

561 QM-240 2438828.0 14691051.0 4884.1 180 50 550 QM_2011 

562 QM-241 2438586.0 14691050.0 4858.1 180 55 500 QM_2011 

563 QM-242 2438083.0 14691081.0 4845.7 180 50 500 QM_2011 

564 QM-243 2437579.2 14691070.0 4825.6 180 45 400 QM_2011 

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565 QM-244 2437352.8 14691071.0 4857.2 180 45 670 QM_2011 

566 QM-245 2437587.0 14691563.0 4897.5 180 45 125 QM_2011 

567 QM-246 2437334.5 14691533.0 4914.5 180 45 140 QM_2011 

568 QM-247 2437074.5 14691016.0 4910 0 90 360 QM_2011 

569 QM-248 2436579.5 14691088.0 4952 180 70 400 QM_2011 

570 QM-249 2436086.5 14691079.0 4924.6 180 60 450 QM_2011 

571 QM-250 2440064.2 14689046.0 4651.9 0 90 400 QM_2011 

572 QM-251 2439809.5 14689069.0 4646.8 0 45 500 QM_2011 

573 QM-252 2440303.8 14689577.0 4622 0 90 400 QM_2011 

574 QM-253 2440319.0 14689046.0 4657.9 0 90 400 QM_2011 

575 QM-254 2439829.2 14688560.0 4650 0 90 400 QM_2011 

576 QM-255 2438575.8 14689545.0 4715.7 0 90 400 QM_2011 

577 QM-256 2438584.8 14690056.0 4734.6 180 45 500 QM_2011 

578 QM-257 2437345.8 14689562.0 4798.8 0 90 500 QM_2011 

579 QM-258 2436821.8 14688528.0 4864.2 0 90 450 QM_2011 

580 QM-259 2436323.5 14688557.0 4921.8 180 45 445 QM_2011 

581 QM-260 2435869.2 14691005.0 4942.7 90 45 500 QM_2011 

582 QM-261 2435860.5 14691014.0 4943.3 0 45 500 QM_2011 

583 QM-262 2435593.0 14691074.0 4911.6 180 45 400 QM_2011 

584 QM-263 2435346.5 14691576.0 4912 180 45 450 QM_2011 

585 QM-264 2440324.5 14690541.0 4695.3 180 55 400 QM_2011 

586 QM-265 2439825.2 14690545.0 4775.2 180 55 500 QM_2011 

587 QM-266 2435599.0 14691584.0 4866.1 180 45 450 QM_2011 

588 QM-267 2436101.5 14691598.0 4863.8 180 50 450 QM_2011 

589 QM-268 2433304.8 14686833.0 5200.4 180 60 300 QM_2011 

590 QM-269 2433067.0 14686832.0 5239.4 0 90 350 QM_2011 

591 QM-270 2435377.5 14690648.0 4956.9 180 45 400 QM_2011 

592 QM-271 2435370.2 14690658.0 4956.7 90 45 500 QM_2011 

593 QM-272 2435805.2 14690551.0 4893.9 180 45 300 QM_2011 

594 QM-273 2439074.8 14689576.0 4689.5 180 60 430 QM_2011 

595 QM-274 2438603.2 14692045.0 4845.6 0 90 600 QM_2011 

596 QM-275 2437947.8 14689560.0 4751.3 180 45 550 QM_2011 

597 QM-276 2432824.5 14686843.0 5251 0 90 300 QM_2011 

598 QM-277 2432569.0 14686858.0 5255 0 90 250 QM_2011 

599 QM-278 2432800.2 14687351.0 5301.7 0 90 300 QM_2011 

600 QM-279 2432569.5 14687087.0 5297.1 0 90 400 QM_2011 

601 QM-280 2439290.5 14688560.0 4668.8 0 90 400 QM_2011 

602 QM-281 2439063.5 14688552.0 4676.7 0 90 480 QM_2011 

603 QM-282 2439364.5 14688668.0 4666.2 0 45 450 QM_2011 

604 QM-283 2438951.0 14688257.0 4690 180 45 595 QM_2011 

605 QM-284 2438324.0 14688305.0 4713 0 90 450 QM_2011 

606 QM-285 2437801.8 14687943.0 4773.6 0 90 400 QM_2011 

607 QM-286 2437760.0 14688261.0 4738 0 90 460 QM_2011 

608 QM-287 2437579.0 14688064.0 4788.8 0 90 400 QM_2011 

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609 QM-288 2437533.8 14688316.0 4760 0 90 600 QM_2011 

613 QM-292 2436773.0 14687908.0 4809.4 0 45 500 QM_2011 

614 QM-293 2436770.8 14687892.0 4810.4 180 45 500 QM_2011 

615 QM-294 2436757.0 14687934.0 4810.6 90 45 500 QM_2011 

616 QM-295 2439317.5 14688028.0 4655.1 0 45 400 QM_2011 

617 QM-296 2438825.5 14687562.0 4681.2 0 45 370 QM_2011 

618 QM-297 2435815.0 14687572.0 4907.9 0 90 505 QM_2011 

619 QM-298 2434319.8 14687335.0 5061.2 0 90 340 QM_2011 

620 QM-299 2434560.5 14687332.0 5042.9 0 90 300 QM_2011 

621 QM-300 2437311.8 14689060.0 4827.9 0 90 490 QM_2011 

622 QM-301 2438845.0 14692049.0 4841.2 0 50 700 QM_2011 

623 QM-302 2439281.0 14691507.0 4783.5 0 90 500 QM_2011 

624 QM-303 2439838.2 14691958.0 4755.8 270 60 500 QM_2011 

625 QM-304 2438075.0 14692107.0 4915.9 180 60 800 QM_2011 

626 QM-305 2437582.0 14691264.0 4839 0 90 500 QM_2011 

627 QM-306 2437340.0 14691274.0 4851 0 90 500 QM_2011 

628 QM-307 2433570.2 14686841.0 5158.6 0 90 300 QM_2011 

629 QM-308 2434065.0 14687079.0 5146.2 270 45 250 QM_2011 

630 QM-309 2433823.2 14686849.0 5163.7 0 90 300 QM_2011 

631 QM-310 2434582.0 14687082.0 5109 0 90 250 QM_2011 

632 QM-311 2433578.2 14687073.0 5142.7 0 90 300 QM_2011 

633 QM-312 2434308.8 14686842.0 5133.4 0 90 200 QM_2011 

634 QM-313 2439082.5 14691046.0 4862.2 0 90 400 QM_2011 

635 QMC-1aR 2438823.2 14690864.0 4823.4 0 90 340 Quaterra-RC-2009 

636 QMC-1bR 2438819.5 14690856.0 4835.8 270 45 450 Quaterra-RC-2009 

637 QMC-21R 2438850.2 14691573.0 4819.4 0 90 400 Quaterra-RC-2009 









646 QMC-4bR 2439271.0 14691489.0 4778.8 270 45 400 Quaterra-RC-2009 

647 QMCC-1 2438827.5 14690867.0 4836.4 0 90 404 Quaterra-Core-2009 







654 QMCC-16 2435832.5 14690068.0 4896 0 90 325 Quaterra-Core-2009 

655 QMCC-17 2435315.8 14689632.0 5012.6 0 90 327.5 Quaterra-Core-2009 

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667 QME-1 2439631.8 14689935.0 4662 0 90 324 Quaterra-Core-2009 

668 QME-10R 2440620.8 14691016.0 4659.3 0 90 400 Quaterra-RC-2009 

669 QME-2 2439495.8 14689994.0 4668.4 0 90 300.5 Quaterra-Core-2009 

670 QME-3 2438029.5 14687909.0 4736.1 0 90 303 Quaterra-Core-2009 

671 QME-4 2437379.5 14687894.0 4830 0 90 115 Quaterra-Core-2009 

672 QME-4aR 2437360.8 14687885.0 4830.5 0 90 230 Quaterra-RC-2009 

673 QME-5 2438309.0 14687637.0 4729.7 0 90 72.5 Quaterra-Core-2009 

674 QME-5aR 2438310.2 14687629.0 4730.1 210 50 80 Quaterra-RC-2009 




678 QME-77R 2442063.2 14689678.0 4563 0 90 350 Quaterra-RC-2009 







685 QMT-1 2437253.8 14688737.0 4810.5 0 90 300 QM-Core-Twin-2010 

686 QMT-10 2439185.8 14689513.0 4680.6 0 90 480 QM-RC-Twin-2010 

687 QMT-10aR 2439180.2 14689554.0 4682.1 30 55 350 QM-RC-Twin-2010 

688 QMT-10bR 2439175.8 14689545.0 4682.4 0 90 350 QM-RC-Twin-2010 




692 QMT-12aR 2438239.0 14687972.0 4715 0 90 110 QM-RC-Twin-2010 

693 QMT-13 2439285.2 14689237.0 4610.2 0 90 309.2 QM-Core-Twin-2010 







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707 QMT-18bR 2440204.2 14689653.0 4630 0 90 350 QM-RC-Twin-2010 



710 QMT-1bR 2437268.5 14688741.0 4810 0 90 300 QM-RC-Twin-2010 

711 QMT-2 2437253.8 14688729.0 4810 210 55 300 QM-Core-Twin-2010 













* RC mean Reverse Circulation 

ⱡ record numbers are not sequential because 61 records were excluded 

from model 

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