preliminary economic assessment 25,000 tpd mill with an … · 2018-12-13 · redhawk copper, inc....

!

REDHAWK COPPER, INC. Copper Creek Project

Preliminary Economic Assessment 25,000 TPD Mill with an Underground Mine

for Development of the Copper Creek Resource

Prepared For:

Redhawk Copper, Inc. P.O. Box 10

San Manuel, Arizona 85631-0010

PREPARED BY:

Mr. Joseph M. Keane, P.E. Mr. Herb Welhener, MMSA-QPM

Mr. Steve Milne, P.E. Mr. Gene Muller, P.E. Mr. David Nicholas

SGS Metcon/KD Engineering 7701 N. Business Park Drive

Tucson, AZ 85743

Document No. Q400-05-028 SGS Project No. 400-05

25 July 2013 AMENDED 28 October 2013

Redhawk Copper, Inc. - Copper Creek PEA 25,000 TPD MillTABLE OF CONTENTS

SECTION PAGE

1.0 SUMMARY 7 ...................................................................................................

2.0 INTRODUCTION 27 ..........................................................................................

3.0 RELIANCE ON OTHER EXPERTS 30 .............................................................

4.0 PROPERTY DESCRIPTION AND LOCATION 31 ............................................

5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY 35 ........................

6.0 HISTORY 36 ......................................................................................................

7.0 GEOLOGICAL SETTING AND MINERALIZATION 46 ....................................

8.0 DEPOSIT TYPES 50 .........................................................................................

9.0 EXPLORATION 51 ............................................................................................

10.0 DRILLING 54 ....................................................................................................

11.0 SAMPLE PREPARTION, ANALYSES AND SECURITY 57 .............................

12.0 DATA VERIFICATION 61 ..................................................................................

13.0 MINERAL PROCESSING AND METALLURGICAL TESTING 86 ...................

14.0 MINERAL RESOURCE ESTIMATES 104 ..........................................................

15.0 MINERAL RESERVE ESTIMATES 138 .............................................................

16.0 MINING METHODS 139 .....................................................................................

17.0 RECOVERY METHODS 161 ..............................................................................

18.0 INFRASTRUCTURE 167 ....................................................................................

19.0 MARKET STUDIES AND CONTRACTS 185 .....................................................

Document No. Q400-05-028 25 July 2013

SGS FORM No. A263a-2/14/13 !

Redhawk Copper, Inc. - Copper Creek PEA 25,000 TPD MillTABLE OF CONTENTS

SECTION PAGE

20.0 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT 186 ....................................................

21.0 CAPITAL AND OPERATING COSTS 204 ..........................................................

22.0 ECONOMIC ANALYSIS 230 ...............................................................................

23.0 ADJACENT PROPERTIES 241 .........................................................................

24.0 OTHER RELEVANT DATA AND INFORMATION 242 ......................................

25.0 INTERPRETATION AND CONCLUSIONS 243 ..................................................

26.0 RECOMMENDATIONS 244 ...............................................................................

27.0 REFERENCES 250 ............................................................................................

28.0 APPENDICES 252 .............................................................................................

Appendix 1 – Redhawk Copper, (USA), Inc. Property List Appendix 2 – CNI Report Appendix 3 – Hydraulic Fill Plant Process Flowsheet Appendix 4 – Process Plant Design Criteria Appendix 5 – Process Plant Flowsheets Appendix 6 – Process Plant Equipment List Appendix 7 – A Siting Study for Production Well Locations B State Land Well Aquifer Testing C Well Spacing Study Appendix 8 – Process Plant Capital Cost Estimate


SGS FORM No. A263a-2/14/13 !

Redhawk Copper, Inc. - Copper Creek PEA 25,000 TPD Mill Page ! 1

List of Tables

Table 1.1 Post-Pillar Cut and Fill Mineralized Material 9 .......................................................................Table 1.2 Diluted/Recovered Post-Pillar Cut and Fill Mineralized Material 10 ........................................Table 1.3 Cu-Mo Second Cleaner Flotation on Composite Samples 15 .................................................Table 1.4 Mine Capital Cost Summary 21 ...............................................................................................Table 1.5 Summary Process Plant – Capital Cost for Base Case 25,000 TPD 22 .................................Table 1.6 Summary of Total Project Capital Costs 23 .............................................................................Table 1.7 Full Production Mine Operating Cost Summary 24 .................................................................Table 1.8 Summary of Plant Operating Costs by Cost Item 24 ...............................................................Table 1.9 Summary of Key Financial Parameters 25 ..............................................................................

Table 2.1 Summary of Qualified Persons 28 ...........................................................................................

Table 10.1 Drill Hole Statistics by Company 54 ......................................................................................Table 10.2 Drilling Statistics by Drilling Type 55 .....................................................................................Table 10.3 Drilling Statistics by Mineralized Area 55 ..............................................................................

Table 12.1 Assays on Copper Standards 62 ...........................................................................................Table 12.2 Assays on Molybdenum Standards 63 ..................................................................................Table 12.3 ALS Database Assays (on Jacobs Pulps) Vs Check Assays for Copper 69 .........................Table 12.4 ALS Database Assays (on Jacobs Pulps) Vs Check Assays for Molybdenum 70 .................Table 12.5 ALS Database Assays (on Jacobs Pulps) Vs Check Assays for Silver 70 ............................

Table 13.1 Effect of Grind on Copper and Molybdenum Recovery in MSRDI Tests 87 ..........................Table 13.2 MSRDI Concentrate Cleaning Summary 87 ..........................................................................Table 13.3 MSRDI Locked Cycle Test Results 88 ...................................................................................Table 13.4 METCON Composite Test Head Analysis 89 ........................................................................Table 13.5 METCON Rougher Flotation Test Results 89 ........................................................................Table 13.6 Summary Results of Copper-Moly Separation 90 .................................................................Table 13.7 METCON Copper – Molybdenum Separation Test Results 91 .............................................Table 13.8 Concentrate Analysis 94 ........................................................................................................Table 13.9 Head Assays Summary of Results 95 ...................................................................................Table 13.10 Cu-Mo Second Cleaner Flotation on Composite Samples 96 .............................................Table 13.11 Head Assays on Composite Samples 97 ............................................................................Table 13.12 Molybdenum Fifth Cleaner Flotation Open Pit Composite Sample 98 ................................Table 13.13 Molybdenum Cleaner Flotation Kinetics Underground Composite Sample 99 ...................Table 13.14 Bond Crusher Work Index 100 ..............................................................................................Table 13.15 Bond Rod Mix Work Index 100 ..............................................................................................Table 13.16 Bond Ball Mill Work Index 100 ..............................................................................................Table 13.17 Abrasion Index 101 ................................................................................................................Table 13.18 JKTech Drop-weight and SMC Test Results 101 ...................................................................Table 13.19 Process Plant Design Parameter 102 ...................................................................................

Table 14.1 December 2012 Mineral Resource Estimate 106 ....................................................................Table 14.2 Drilling and Assaying Statistics, All Data 106 ..........................................................................Table 14.3 Drilling Statistics by Mineralized Area 107 ..............................................................................Table 14.4 Composite Statistics by Zone Boundaries 108 ........................................................................Table 14.5 Assay and 20ft Bench Composite Statistics, All Data 108 .......................................................Table 14.6 20ft Bench Composites Statistics by Grade Zones 109 ..........................................................Table 14.7 Covariance Copper Variograms, 20ft Composites 112 ...........................................................Table 14.8 Copper Creek December 2012 Mineral Resource Estimate 134 ............................................Table 14.9 Individual Breccia Deposits at 0.75% CuEq Cutoff 135 ............................................................Table 14.10 Keel-American Eagle Mineral Resource 136 ........................................................................

List of Tables (continued)

Table 14.11 Acid Soluble to Total Copper Ratio, Total Copper 137 ...........................................................

!


Table 16.1 Post-Pillar Cut and Fill Mineralized Material 142 .....................................................................Table 16.2 Diluted/Recovered Post-Pillar Cut and Fill Mineralized Material 143 ......................................Table 16.3 Preproduction Development Quantities and Construction Time Estimates 150 ......................Table 16.4 Preproduction Personnel 151 ..................................................................................................Table 16.5 Personnel Summary 154 .........................................................................................................Table 16.6 Estimated Service Water Usage 157 .......................................................................................Table 16.7 U.G. Power Requirements 158 ...............................................................................................Table 16.8 Ventilation Requirements 159 .................................................................................................Table 16.9 Redhawk Resources 160 ........................................................................................................

Table 18.1 Water Balance Calculations Backfill Plant Rejects Thickened 169 .........................................Table 18.2 Summary of Water Uses, Copper Flat Project, 25,000 tpd Mining Rate 176 ..........................Table 18.3 Height Versus Capacity Relationships 179 .............................................................................

Table 20.1 Hydrogeological Cost Estimate 189 ........................................................................................Table 20.2 Environmental Permitting Analysis for Copper Creek Project 202 ..........................................

Table 21.1 Copper Creek PEA Mine Capital Cost Summary 204 .............................................................Table 21.2 Estimated Annual Capital 205 .................................................................................................Table 21.3 Hydraulic Backfill Plant Capital Costs Estimate (±40%) 207 ...................................................Table 21.4 Hydraulic Backfill Plant Operating Cost Estimate (±40%) 207 ................................................Table 21.5 Full Production Mine Operating Cost Summary 208 ...............................................................Table 21.6 Estimated U.G. Mine Annual Operating Expense 209 ............................................................Table 21.7 Labor Basis – Operating Labor Cost Estimate 210 .................................................................Table 21.8 Post Pillar Cut and Fill PEA Study 211 ....................................................................................Table 21.9 G&A Costs 212 ........................................................................................................................Table 21.10 Summary Process Plant – Capital Cost for Base Case 25,000 TPD 213 .............................Table 21.11 Summary of Plant Operating Costs by Cost Item 219 ...........................................................Table 21.12 Power Consumption Summary 219 .......................................................................................Table 21.13 Power Costs 219 ...................................................................................................................Table 21.14 Labor Costs 220 ....................................................................................................................Table 21.15 Reagent Costs 221 ................................................................................................................Table 21.16 Grinding Media Operating Cost Estimates 222 .....................................................................Table 21.17 Wear Material Operating Cost Estimates 222 .......................................................................Table 21.18 Water Cost Estimate 223 .......................................................................................................Table 21.19 Repair Minerals and Operating Supplies 223 ........................................................................Table 21.20 Tailings Storage Facility Capital Costs 225 ...........................................................................Table 21.21 Diversion Tunnel Cost Estimate 227 .....................................................................................Table 21.22 Tailings Storage Facility Closure and Reclamation Costs 229 ..............................................

Table 22.0 Base Case Economic Model Metal Price Assumptions 230 ....................................................Table 22.1 Mine Production Schedule 234 ...............................................................................................Table 22.2 Mining Schedule 234 ...............................................................................................................Table 22.3 Flotation Concentrator Filed Production Schedule 234 ...........................................................Table 22.4 Flotation Concentrate Production Schedule 235 .....................................................................Table 22.5 Concentrate Smelted Schedule 235 ........................................................................................Table 22.6 Payable Metal and Commodity Schedule 235 ........................................................................Table 22.7 Net Smelter Return Schedule 236 ...........................................................................................Table 22.8 Cash Operating Cost Schedule 237 ........................................................................................Table 22.9 Cash Operating Cost per Ton Mined Schedule 238 ................................................................

List of Tables (continued)

Table 22.10 Byproduct Credit Schedule 238 .............................................................................................Table 22.11 Cash Operating Calculation Schedule 239 ............................................................................Table 22.12 Capital Cost Schedule 239 ....................................................................................................Table 22.13 Cash Flow Schedule 240 ......................................................................................................

Table 26.1 Drilling and Installation and Testing of the Redhawk Minimal Test Well 247 ...........................

!


!


List of Figures

Figure 1.1 Location Map 7 ....................................................................................................................

Figure 4.1 Location Map 32 ....................................................................................................................Figure 4.2 Arizona Copper Infrastructure 33 ...........................................................................................

Figure 7.1 Regional Geology Map 47 .....................................................................................................Figure 7.2 Local Geology – Mapped Lithologies 48 ................................................................................

Figure 10.1 Drill Hole Locations 56 .........................................................................................................

Figure 12.1 Standard 1, Copper, Expected Value 0.042% 64 ................................................................Figure 12.2 Standard 2, Copper, Expected Value 0.969% 64 ................................................................Figure 12.3 Standard 3, Copper, Expected Value 0.555% 65 ................................................................Figure 12.4 Standard 4, Copper, Expected Value 5.678% 65 ................................................................Figure 12.5 Standard 5, Copper, Expected Value 0.014% 66 ................................................................Figure 12.6 Standard 1, Molybdenum, Expected Value 0.0030% 66 .....................................................Figure 12.7 Standard 2, Molybdenum, Expected Value 0.0075% 67 .....................................................Figure 12.8 Standard 3, Molybdenum, Expected Value 0.0092% 67 .....................................................Figure 12.9 Standard 4, Molybdenum, Expected Value 0.0010% 68 .....................................................Figure 12.10 Standard 5, Molybdenum, Expected Value 0.0003% 68 ...................................................Figure 12.11 ALS Database Assays vs. Inspectorate Check Assays, Copper (XY) 71 ...........................Figure 12.12 ALS Database Assays vs. ALS Assays on Second Pulp, Copper (XY) 72 ........................Figure 12.13 ALS Database Assays vs. METCON AA Check Assays, Copper (XY) 73 .........................Figure 12.14 ALS Database Assays vs. METCON AA2 Check Assays, Copper (XY) 74 .......................Figure 12.15 ALS Database Assays vs. METCON ICP Check Assays, Copper (XY) 75 ........................Figure 12.16 ALS Database Assays vs. Inspectorate Check Assays, Copper (QQ) 76 ..........................Figure 12.17 ALS Database Assays vs. ALS Assays on Second Pulp, Copper (QQ) 77 ........................Figure 12.18 ALS Database Assays vs. Inspectorate Check Assays, Molybdenum (XY) 78 ..................Figure 12.19 ALS Database Assays vs. ALS Duplicate Assays, Molybdenum (XY) 79 ..........................Figure 12.20 ALS Database Assays vs. METCON AA Check Assays, Molybdenum (XY) 80 ................Figure 12.21 ALS Database Assays vs. METCON AA2 Check Assays, Molybdenum (XY) 81 ..............Figure 12.22 ALS Database Assays vs. METCON ICP Check Assays, Molybdenum (XY) 82 ...............Figure 12.23 ALS Database Assays vs. Inspectorate check Assays, Silver (XY) 83 ..............................Figure 12.24 ALS Database Assays vs. ALS Assays on Second Pulp, Silver (XY) 84 ...........................Figure 12.25 ALS Database Assays vs. METCON Check Assays, Silver 85 ..........................................

Figure 13.1 MSRDI Rougher Flotation Kinetics 88 .................................................................................Figure 13.2 Redhawk 2007 Drill Results 94 ............................................................................................Figure 13.3 Redhawk Summary Flowsheet 103 .......................................................................................

Figure 14.1 Drill Hole Locations 107 .........................................................................................................Figure 14.2 20ft. Composite Cumulative Frequency, Copper and Molybdenum, All Data 109 .................Figure 14.3 20ft. Composite Cumulative Frequency, Copper by Zone 110 ..............................................Figure 14.4 20ft. Composite Cumulative Frequency, Molybdenum by Zone 111 .....................................Figure 14.5 Omnidirectional Covariance Copper Variogram, 20ft Copper Composites, All Zones 113 ....Figure 14.6 20ft. Composite Grade, American Eagle 114 ........................................................................Figure 14.7 20ft. Composite Grade, Mammoth-Keel and American Eagle 115 ........................................Figure 14.8 Model Block Grad Distributions 117 ......................................................................................Figure 14.9 Grade Zone Boundaries with Composites on the 1980 Bench 118 .......................................Figure 14.10 Copper Block Grade, American Eagle 120 ..........................................................................Figure 14.11 Copper Block Grade, Keel – American Eagle 121 ...............................................................Figure 14.12 Copper Block Grades, Bench 1980 122 ..............................................................................

List of Figures (continued)

Figure 14.13 Grade Thickness Copper in Model Blocks 123 ....................................................................Figure 14.14 Grade-Thickness Molybdenum in Model Blocks 124 ...........................................................

!


Figure 14.15 Kriging Variance Vs. Number of Holes in Search Ellipsoid 125 ...........................................Figure 14.15 Kriging Variance Vs. Number or Holes in Search Ellipsoid (Continued) 126 .......................Figure 14.16 Resource Classification, American Eagle 128 .....................................................................Figure 14.17 Resource Classification, Mammoth-Keel and American Eagle 129 .....................................Figure 14.18 Grade-Thickness Copper, Feet-Percent, Measure and Indicated Blocks 130 .....................Figure 14.19 Grade-Thickness Copper, Feet-Percent, Inferred Blocks 131 .............................................

Figure 16.1 Deposit with 70 Degree Crack Limits 145 ..............................................................................Figure 16.2 Mining Method Schematic 148 ..............................................................................................Figure 16.3 Typical Level / Shaft Section 149 ...........................................................................................Figure 16.4 25,000 tpd Preproduction Development Schedule 152 .........................................................

Figure 18.1 Tailings Storage Facility 177 ..................................................................................................Figure 18.2 Hydraulic Backfill Plant Process Flowsheet 182 ....................................................................

Figure 22.1 IRR Sensitivity (at US$3.00 per lb. Cu) 232 ..........................................................................Figure 22.2 NPV Sensitivity (at US$ 3.00 per lb. Cu) 233 ........................................................................

!


Glossary

Below Ground Surface bgs ................................................................................. Day d ........................................................................................................................................ Days per year (annum) dpa Degree ◦ Foot / Feet ft Feet above sea level fasl Grams per ton gpt Greater than > Horsepower hp Hour h Hours per day h/d Less than < Life of Mine LOM Metric ton tonne Million M Million tons Mt Million tons per year (annum) Mtpa Minute (plane angle) ‘ Minute (time) min Ounce oz Percent % Pound lb Second (time) s Short Ton t Tons per day tpd Tons per hour tph Tons per year tpa Year (annum) a

!


1.0 SUMMARY

Following is a summary of the relevant details of the Copper Creek Project. Supplemental information is presented in the report text and appendices.

1.1 Property Description and Location

The Copper Creek project is located about 75 road miles northeast of Tucson, Arizona USA. The property lays approximately 10 miles from the town of Mammoth, in the Bunker Hill Mining District on the western slopes of the Galiuro Mountains, Pinal County, Arizona, USA (indicated by red star). It lies in Township 8S, Range 18E, Sections 1, 2, 3, 4, 9, 10, 11, 12, 13, 14 and 15 GSRBM and is centered at 32°45' N Latitude, 110°30' W Longitude. (Figure1.1). The property is owned by Redhawk Copper, USA a wholly owned subsidiary of Redhawk Resources Inc. of Vancouver, BC Canada.

!

Figure 1.1 - Location Map

1.2 Geological Setting and Deposit Types

In addition to numerous breccia pipe deposits, the Copper Creek property is host to the American Eagle/Keel (AEK) “early-halo” type copper-molybdenum porphyry deposits, at depth. The Keel is located below the Mammoth breccia, and the AE is the continuation of the porphyry deposit situated to the east and south of the Keel. Above

!


the porphyry system, over three hundred breccia occurrences, overlying the AEK, and extending to the north, northwest, west, and southeast have been identified. Previous studies have evaluated the potential of mining the Childs, Mammoth, Old Reliable, Prince and Globe breccia deposits by selective mining methods.

At depth, where drilling density is sufficient, the base of the breccia pipes neck down into small gray or dark porphyry bodies. These appear to be the feeder zone to the breccia mineralization. The Keel zone (a zone of sheeted veins, veinlets and stockworks) is situated below the Mammoth breccia, while the AE porphyry is situated below several breccias that outcrop.

The AEK deposit exhibits disseminated mineralization in the form of clots, low angled, "sheeted" early dark micaceous (EDM) vein sets, high angled EDM vein sets, well mineralized intense quartz-sericite zones, and some areas that are likely deep-seated breccias.

Based on detailed core logging and surface mapping, the AEK deposit is described as a sheeted vein, early-halo type, porphyry-like system (cf. Proffett, 2009). The mineralization is predominantly structurally controlled by fracture systems of several orientations, with some mineralization being disseminated. The mineralization is zoned with depth, going from predominately pyrite, near surface, and transitioning to chalcopyrite, and then to bornite, with depth.

At Copper Creek, copper oxide minerals have very limited replacement of the sulfides in the near-surface supergene weathering environment. A chalcocite and copper oxide enriched, supergene zone forms the principal part of the Old Reliable deposit above the water table. Molybdenite occurs in sufficient amounts in the AEK porphyry, as well as the Childs-Aldwinkle breccia, and the Old Reliable breccia, to be of interest economically. Gold and silver occur occasionally and have not been systematically assayed by previous owners. Redhawk has re-assayed individual intervals to better understand the precious metal grade distribution, where previous owners had assayed long composite runs.

Based on recent core drilling in the AEK, a significant tonnage of indicated and inferred mineralization has been identified. Based on these results, studies have been initiated to determine the economic potential of mining these deposits by bulk underground mining methods, or possibly by an open pit.

The existing models for the AEK deposit were updated with the recent drill hole information by Independent Mining Consultants (IMC) of Tucson, Arizona, using blocks 100 ft by 100 ft by 25 ft high. From the model, mineralization limits at various cutoff grades, between the surface and depths of almost 4,000 ft, were determined.

!

! 1

1.3 Mineral Resource Estimates

The resource tons and grades used for the PEA mine plan are contained within the American Eagle, Keel, Mammoth and lower Childs deposit grade shells, at a breakeven cutoff grade of 0.55 percent Cueq, and were tabulated by IMC, for each 20 ft bench height, between elevations 500 and 4200, with the results presented in the following Table 1.1. This preliminary economic assessment (PEA) is preliminary in nature and includes inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves. There is no certainty that the preliminary economic assessment will be realized. Mineral resources that are not mineral reserves do not have demonstrated economic viability.

A new 20 ft x 20 ft x 20 ft model was developed by IMC based on the latest drilling input, and new tonnages and grades tabulated. Bench plans on 20 ft centers were drawn to indicate above cutoff areas in the Keel and AE deposits.

Using MineSight software, four areas, above the 0.55 percent Cueq cutoff grade, were then identified, in the AE, Keel, Mammoth and lower Childs deposits, that contained a sufficient tonnage, which could be mined by a post-pillar cut and fill mining method, and each could support a production rate of 6,250 t/d. Haulage level elevations were then located at the base of each area.

Table 1.1 Post-Pillar Cut & Fill Mineralized Material

Classification Deposit Tons Grade (%Cueq.)

Contained Klbs Cu

Indicated American Eagle 84,364,000 0.76 1,282,352

Indicated Keel 40,665,000 0.75 609,975

Measured/Indicated Mammoth 4,547,000 1.32 120,041

Indicated Childs Prince/Globe 2,464,000 1.66 81,805

Total 132,040,000 0.79 2,094,173

Inferred American Eagle 55,024,000 0.74 814,355

Inferred Keel 17,276,000 0.66 228,043

Inferred Mammoth 527,000 1.03 10,856

Inferred Childs Prince/Globe 1,244,000 1.60 39,808

Total 74,071,000 0.74 1,093,062

Document No. Q400-05-028 25 July 2013 Amended 28 October 2013

SGS FORM No. A263a-2/14/13 !

! 2

The individual bench maps within the deposit grade shells were then reviewed for above cutoff grade blocks that would not support development access, or needed to be re-classified by location. These changes were then tabulated to reflect the minable tons and grades available within the resource shells. Factors for mining (external) dilution and recovery, based on the proposed mining method, were applied. The results of these exercises indicate the potentially minable tons and grades within the resource grade shells for the various deposits, and were used in the cashflow evaluation of the Copper Creek deposits.

The following Table 1.2 provides a summary of the diluted/recovered resources which make up the PEA mine schedule.

The mine plan of this (PEA) is preliminary in nature and includes inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves. There is no certainty that the preliminary economic assessment will be realized.

1.4 Mining

Table 1.2 Diluted/Recovered Post-Pillar Cut & Fill Mineralized Material*

Classification & Deposit KTonsGr.

(%Cu)Gr.

(%Mo) Ag (oz/t) KLbs Cu

Indicated

American Eagle 63,905 0.76 0.012 0.03 971,356

Keel 30,804 0.75 0.025 0.12 462,060

Mammoth 3,026 1.27 0.003 0.12 76,860

Chiles/Prince/Globe 2,013 1.65 0.046 0.13 66,832

Total 99,748 0.790 0.016 0.063 1,577,108

Inferred

American Eagle 41,681 0.74 0.012 0.03 603,031

Keel 13,086 0.66 0.018 0.10 167,179

Mammoth 338 1.03 0.004 0.10 7,202

Childs/Prince/Globe 1,037 1.40 0.021 0.08 28,969

Total 56,142 0.735 0.014 0.048 806,381

*Extensions may vary due to rounding


SGS FORM No. A263a-2/14/13 !

! 3

The basic mine plan would access the Mammoth-Keel and AE deposits mine through a 33 ft diameter, 3,800 ft deep, vertical shaft located at coordinates 638765 N and 401629 E. The current location was the result of evaluating four different locations, which considered the surface road access, deposit locations, Redhawk “patented” property boundaries, Old Reliable pit limits, and the projected geotechnical crack lines to the surface. This shaft size is capable of handling a production rate of 25,000 t/d.

Four major production levels (16 ft x 16 ft in Section) were selected. These levels would allow access from the shaft to the base of the most favorably-mineralized areas in the Mammoth, Lower Childs, Keel and AE deposits. The uppermost level (3000 Level) would access the bottom on the Mammoth and Lower Childs deposits,


SGS FORM No. A263a-2/14/13 !

! 1

while the 1000 and 1500 Levels would access both the Keel and AE deposits. The lowest level (500 Level) would access the deeper extensions of the Keel deposit only.

Internal ramps (15 ft x 15 ft), from the major haulage levels, would allow access to successive mining lifts and the base of the secondary deposits located above the haulage levels. The initial production would be situated in four stoping areas - two in the Keel (one above the 1000 Level, and one above the 1500 Level), and two in the AE deposit - (one above the 1000 Level, and one above the 1500 Level). Each stope would be designed for a production rate of 6,250 tons per day, for a total production rate of 25,000 tons per day.

Two, 20 ft diameter, approximately 3,000 ft deep, ventilation shafts would be sunk from the surface to the lowest level in each of the Mammoth-Keel and AE deposits, with connections to each of the haulage levels. Ten-ft diameter, bored passes would connect the mining levels with the haulage levels.

Mining would proceed from the haulage level upwards in 20 ft lifts, utilizing a post-pillar cut and fill mining method. Blast-hole drilling would utilize electric-hydraulic, two-boom, jumbos. Loading would be performed with 9 to 10 cy LHD’s, which would move the blasted rock from the faces, or muck bays, to passes that would transfer the blasted rock to the haulage level. Fifty-ton haulage trucks would move the rock from the passes to grizzlies, equipped with a rock breaker, located on each haulage level, approximately 1,300 ft from the shaft. The broken rock would pass through the grizzlies to 60-in jaw crushers, be reduced to a nominal 6-in product, and then be moved via an inclined conveyor to a 4,000 ton storage pocket located adjacent to the production shaft. Fifty-five-ton skips would move the rock through the shaft to the surface.

When a 20 ft lift has been exploited, bulkheads will be installed to contain an engineered backfill consisting of deslimed mill tailings, classified mine waste, fly ash, and cement. When this backfill has consolidated, the next lift can be started on the top of the previously-excavated lift. The use of an engineered backfill will allow smaller pillars than would normally be allowed, to be left. This will optimize mining recovery of material from each of the deposits. In addition, by returning a large portion of the mill tailings back into the mine, the surface tailings impoundment size, and cost, can be minimized.

In general, the rocks at Copper Creek are very strong, with an average compressive strength of 10,000 to 20,000 psi, and median RQD’s in the range of 80 to 90 percent. The phreatic surface has been interpreted to be at the level of the San Pedro river, which is around the 2,000 ft elevation. The surface elevation over the deposits is around the 4,200 ft elevation, with the upper limit of reasonably contiguous mineralization in both deposits being around the 2,000 ft elevation. The better mineralization, in the both deposits, continues downward to approximately the 500 ft elevation. Preliminary geotechnical investigations indicate that rooms 20 ft wide and 20 ft high can be excavated in the rocks in the Mammoth-Keel and AE deposits,


SGS FORM No. A263a-2/14/13 !

! 2

leaving 20 ft square pillars for initial support. This would result in a mining recovery of 75 percent. The RQD’s in the breccias (Mammoth/Childs) are somewhat less, thusly CNI has recommended leaving 25 ft x 25 ft pillars in these areas, resulting in an overall recovery there of 65 percent.

The upper (above the 3500L) Childs, Prince, and Globe breccia deposits would be developed and mined by the previously-proposed decline and access drifts, with a cut and fill mining method. A third ventilation shaft, as previously proposed, would be located adjacent to the upper Childs deposit access.

1.4.1 Geotechnical Evaluation

Call & Nicholas, Inc. (CNI) has evaluated the American Eagle, Keel, Mammoth and Childs deposits as potential post-pillar cut and fill operations (PPCF). The American Eagle and Keel deposits have similar RQD distributions and intact strength properties, which are greater than those of the Mammoth and Childs.

Based on surface geologic structures, a room width of 20 ft is recommended using 4-ft long split set bolts on 3 to 4 ft centers. As additional joint set data from within the deposits are gathered, there is the potential for the American Eagle and Keel to be widened to a 25 ft room width, while in the Mammoth and Childs room widths may need to be reduced.

The recommended pillar widths for the American Eagle and Keel are 20 by 20 ft, and in the Mammoth and Childs 25 by 25 ft. These pillar widths are based on mining 15ft high lifts for five (5) lifts after which mining heights can be increased to 20 ft.

Because the Mammoth and Childs deposits are at or near surface and because the actual mining area is yet to be finalized, CNI used a 70 degree crack angle to identify where Copper Creek is likely to be impacted and where to locate potential shaft locations.

CNI recommends diverting Copper Creek when either of the following happens:

▪ The fill height of the American Eagle or Keel is 600 feet, or ▪ When mining starts at either the Mammoth or the Childs. ▪ When post pillar cut and fill mining starts at either the Mammoth or Childs.

The mine schedule includes open stope mining part of the Childs and Mammoth deposits as the access is developed down to the American Eagle and Keel. This mining will be of a limited area and the backfill will be a cemented fill. Therefore the surface subsidence is unlikely.

A mining recovery factor of 75 percent can be expected in the American Eagle and Keel deposits, and a mining recovery of 65 percent can be expected in the breccia post-pillar room and pillar mining


SGS FORM No. A263a-2/14/13 !

! 3

1.4.2 Basic Mine Plan

The basic mine would access the Mammoth-Keel and AE deposits mine through a 33 ft diameter, 3,800 ft deep, vertical shaft located at coordinates 638765 N and 401629 E. These coordinates may change with better definition of the deposits, surface topography, and Redhawk “patented” property boundaries. The current location was the result of evaluating four different locations, which considered the surface road access, deposit locations, Redhawk “patented” property boundaries, Old Reliable pit limits, and the projected geotechnical crack lines to the surface. This shaft size is capable of handling a production rate of 25,000 tons per day.

Four major production levels (16 ft x 16 ft in Section) were selected. These levels would allow access from the shaft to the base of the most favorably-mineralized areas in the Mammoth, Lower Childs, Keel and AE deposits. The uppermost level (3000 Level) would access the bottom on the Mammoth and Lower Childs deposits, while the 1000 and 1500 Levels would access both the Keel and AE deposits. The lowest level (500 Level) would access the deeper extensions of the Keel deposit only.

All four haulage level shaft stations are planned to be developed during the shaft sinking phase to allow level development on the 500, 1000, 1500, and 3000 levels.

Internal ramps (15 ft x 15 ft), from the major haulage levels, would allow access to successive mining lifts and the base of the secondary deposits located above the haulage levels. The initial production would be situated in four stoping areas – two in the Keel (one above the 1000 Level, and one above the 1500 Level), and two in the AE deposit – (one above the 1000 Level, and one above the 1500 Level). Each stope would be designed for a production rate of 6,250 tons per day, for a total production rate of 25,000 tons per day.

Two, 20 ft diameter, approximately 3,000 ft deep, ventilation shafts would be sunk from the surface to the lowest level in each of the Mammoth-Keel and AE deposits, with connections to each of the haulage levels. Ten foot diameter, bored, rock and ventilation passes would connect the mining levels with the haulage levels.

Mining would proceed from the haulage level upwards in 20 ft lifts, utilizing a post-pillar cut and fill mining method. Blast-hole drilling would utilize electric-hydraulic, two-boom, jumbos. Rock loading would be performed with 9 to10 cy LHD’s, which would move the blasted rock from the faces, or muck bays, to rock passes that would transfer the blasted rock to the haulage level. Fifty-ton haulage trucks would move the rock from the passes to grizzlies, equipped with a rock breaker, located on each haulage level, approximately 1,300 ft from the shaft. The broken rock would pass through the grizzlies to 60-inch jaw crushers, be reduced to a nominal 6-inch product, and then be moved via an inclined conveyor to a 4,000 ton storage pocket located


SGS FORM No. A263a-2/14/13 !

! 4

adjacent to the production shaft. Fifty-five ton skips would move the rock through the shaft to the surface.

When a 20 ft lift has been exploited, bulkheads will be installed to contain an engineered backfill consisting of deslimed mill tailings, classified mine waste, fly ash, and cement. When this backfill has consolidated, the next lift can be started on the top of the previously-excavated lift. The use of an engineered backfill will allow smaller pillars than would normally be allowed, to be left. This will optimize mining recovery of material from each of the deposits. In addition, by returning a large portion of the mill tailings back into the mine, the surface tailings impoundment size, and cost, can be minimized.

In general, the rocks at Copper Creek are very strong, with an average compressive strength of 10,000 to 20,000 psi, and a median RQD’s in the range of 80 to 90 percent. The phreatic surface has been interpreted to be at the level of the San Pedro river, which is around the 2,000 ft elevation. The surface elevation over the deposits is around the 4,200 ft elevation, with the upper limit of reasonably contiguous mineralization in both deposits being around the 2,000 ft elevation. The better mineralization, in the both deposits, continues downward to approximately the 500 ft elevation. Preliminary geotechnical investigations indicate that rooms 20 ft wide and 20 ft high can be excavated in the rocks in the Mammoth-Keel and AE deposits, leaving 20 ft square pillars for initial support. This would result in a mining recovery of 75 percent. The RQD’s in the breccias (Mammoth/Childs) are somewhat less, thusly CNI has recommended leaving 25 ft x 25 ft pillars in these areas, resulting in an overall recovery there of 65 percent.

The upper (above the 3500L) Childs, Prince, and Globe breccias deposits would be developed and mined by the previously-proposed decline and access drifts, with a cut and fill mining method. A third, 16 ft diameter, ventilation shaft, as previously discussed, would be located adjacent to the upper Childs deposit access.

1.5 Mineral Processing and Metallurgical Testing

A conceptual process flowsheet was developed for processing 25,000 tons per day with an overall availability factor of 92.5 percent. The process design incorporates modern equipment where applicable. For example, larger flotation cells have been selected for the rougher flotation with vertical mills for the regrinding circuit. Designs incorporated in this study are considered standard mining industry practice.

Locked cycle flotation tests conducted by Mountain States R&D International, Inc. (MSRDI) on the average grade composites showed copper recovery above 95 percent and copper concentrate grades between 32 and 62 percent. Molybdenum recoveries were proportional to the molybdenum head grade with the high-grade sample giving 94 percent recovery and the low-grade sample giving 28 percent recovery.


SGS FORM No. A263a-2/14/13 !

! 5

In year 2012, METCON Research (METCON) conducted open cycle copper-molybdenum second cleaner flotation testing using the scheme of reagents developed by MSRDI on 14 composite samples representing different zones from the Copper Creek Project. The composite samples were identified as follows:

▪ Composite 1 - Copper Grade in the 0.2 to 0.3 Percent Range ▪ Composite 2 - Chalcopyrite Dominant ≥ 0.2 to 0.5 Percent Copper ▪ Composite 3 - Chalcopyrite Dominant ≥ 0.5 Percent Copper ▪ Composite 4 - Bornite Moderate to Strong ≥ 0.2 to 0.5 Percent Copper ▪ Composite 5 - Bornite Moderate to Strong ≥ 0.5 Percent Copper ▪ Composite 6 - High Copper Grade ▪ Composite 7 - Mid Copper Grade ▪ Composite 8 - Low Copper Grade ▪ Composite 9 - SE Low Grade ▪ Composite 10 - SE Mod-High Grade ▪ Composite 11 - SE High Bornite ▪ Composite 12 - SW Low Copper Grade ▪ Composite 13 - SW Moderate High Copper Grade ▪ Composite 14 - SW High Bornite

The metallurgical data developed on the open cycle copper-molybdenum second cleaner flotation testing are summarized in Table 1.3 below.

Table 1.3 Cu-Mo Second Cleaner Flotation On Composite Samples

Sample ID

Cu-Mo Second Cleaner Concentrate Recovery (%)

Cu (%)

Mo (%)

Au (g/t)

Ag (g/t) Cu Mo Au Ag

Composite 1 - Copper Grade in the 0.2 to 0.3 Percent Range 28.80 0.56 1.20 NA 86.71 75.34 NA NA

Composite 2 - Chalcopyrite Dominant Copper Grade ≥ 0.2 to 0.5 Percent 30.50 0.39 1.40 NA 85.26 72.03 NA NA

Composite 3 - Chalcopyrite Dominant, Copper Grade ≥ 0.5 Percent 30.20 0.75 1.49 NA 87.23 73.76 NA NA

Composite 4 - Bornite Moderate to Strong, Copper Grade ≥ 0.2 to 0.5 Percent 41.80 2.28 3.95 NA 85.43 72.45 NA NA

Composite 5 - Bornite Moderate to Strong, Copper Grade ≥ 0.5 Percent 40.10 0.56 5.66 NA 77.17 80.67 NA NA

Composite 6 - High Copper Grade 31.10 0.20 0.96 NA 88.95 77.40 NA NA

Composite 7 - Mid Copper Grade 23.90 0.20 0.93 NA 87.36 66.46 NA NA

Composite 8 - Low Copper Grade 25.50 0.34 0.95 NA 82.78 65.97 NA NA

Composite 9 - SE Low Copper Grade 18.99 0.04 0.54 47 88.59 37.84 57.37 54.08

Composite 10 - SE Moderate High Grade 21.07 0.16 0.57 61 92.84 80.17 72.42 70.16

Composite 11 - SE High Bornite 21.84 1.07 0.41 56 88.27 87.33 45.47 53.87

Composite 12 - SW Low Copper Grade 20.84 0.79 0.73 46 85.07 86.69 57.74 49.11

Composite 13 - SW Moderate High Copper Grade 31.01 0.03 0.77 44 89.29 38.67 62.83 48.27

Composite 14 - SW High Bornite 31.50 12.30 3.59 154 91.81 97.06 82.40 78.80


SGS FORM No. A263a-2/14/13 !

! 6

Remarks: NA, data not available

The following comments relate to the open cycle copper-molybdenum second

cleaner flotation tests on composite samples from Copper Creek Project.

▪ Copper recovery obtained ranged from approximately 77 percent to 93 percent. The highest copper recovery of approximately 93 percent was observed on Composite 10.

▪ The lowest copper recovery of approximately 77 percent was obtained on Composite 5.

▪ Molybdenum recovery ranged from approximately 38 to 97 percent. Composite 14 provided the highest molybdenum recovery of approximately 97 percent.

▪ The lowest molybdenum recovery of approximately 38 percent was obtained on Composite 9.

▪ Complete mass balances for gold and silver were conducted on Composites 9 through 14.

▪ Additional locked cycle flotation testing should be conducted to determine the ultimate concentrate and tailing values that will be realized when internal products are recycled as occurs in normal plant practice.

1.6 Tailings Storage Facility

The proposed mining operation will process 25,000 tons per day with approximately 50 percent of the process tailings used for underground mine backfill. The tailings storage facility (TSF) will be required to contain approximately 15,750 tons per day of tailings, including fine rejects from the hydraulic backfill operation (cyclone overflow), and whole tailings when the backfill plant is not in operations. A TSF site with a capacity of approximately 90 million tons has been identified on State of Arizona land approximately 4.2 miles west of the proposed Saloon Gulch shaft site. The foundation of the facility is composed on alluvial gravels. As such, borrow materials, for embankment construction, are anticipated to be readily available locally.

TSFs for base metal mining projects in Arizona typically employ an earthen starter dam followed by upstream construction of subsequent raises using tailings sand. Because the TSF will be required to contain fine rejects from the backfill plant, the conventional construction approach is not considered. The preliminary layout of the tailings impoundment assumes a dam constructed with borrowed fill material by the downstream raise method. This downstream construction method can be used to


SGS FORM No. A263a-2/14/13 !

! 7

construct a stable TSF dam regardless of the characteristics of the tailings, however; that method required more embankment fill material. Subject to testing and demonstration of feasibility, the currently proposed design approach could potentially be modified to reduce embankment fill requirements and construction costs.

1.7 Environmental Studies, Permitting and Social or Community Impact

1.7.1 Hydrology and Geochemistry

Building on previous hydrological studies conducted within 1/2-mile of the proposed mine site, Golder Associates conducted a regional hydrogeological investigation and well siting study incorporating major hydrogeological regions of the San Pedro river basin including areas with historic high yield deep production wells.

The lower San Pedro river basin contains two major water-bearing units: the streambed Holocene alluvium that forms the San Pedro River channel and floodplain, and the alluvial basin fill deposits that are composed of a younger basin-fill, older basin-fill, and a basal conglomerate. The basin-fill deposits form the basin’s primary aquifer because of their high permeability and large groundwater storage. However due to recent surface water adjudication, water rights, and permitting issues, groundwater from the upper San Pedro River aquifer including its base flow (Holocene alluvium) should not be considered as potential supply water for the Copper Creek project. The Golder study concentrated on the lower basin fill groundwater that is believed to be hydrologically isolated from the San Pedro River Holocene alluvial system, and identified four locations for potential siting of production wells.

Geochemical analysis of the mine development rock from the Mammoth and Childs-Aldwinkle breccia areas indicate that the country rock in the vicinity of the initial breccia mining operation has a low potential for acid generation, and drainage from the development rock is expected to be of good quality. Similar testing has not been performed for development rock in the vicinity of the deeper porphyry mineralization of the Keel and American Eagle deposits. Future development rock is likely to be consumed in mining backfill operations. Analysis of the geochemical properties of the future tailings has not been performed, though the potential for acid generation and/or leachable arsenic could lead to a requirement for lined tailings disposal facilities to mitigate impacts to water resources.

1.7.2 Environmental Permitting

Environmental permitting for the project will be required at the state and federal level. Some key considerations with regard to cost and schedule include:

▪ Proximity (less than 2 miles) of the Galiuro Wilderness Area, a Class I air shed

▪ Significant historic mining features throughout the project area


SGS FORM No. A263a-2/14/13 !

! 8

▪ Significant NEPA review requirements and associated public involvement

In addition, the project has several favorable attributes from a permitting perspective, including proximity to existing mining districts and associated infrastructure, and remoteness from residential and urban centers.

1.7.2.1 Federal Permitting

Because significant portions of the project are located on lands managed by the Bureau of Land Management (BLM), it is assumed for this evaluation that a Mine Plan of Operations (MPO) will be required for development of the project. Once the MPO has been submitted, the BLM will begin its review process under NEPA. Given the size of the proposed project and recent NEPA evaluations for other mining projects, it is assumed for the purposes of this analysis that an EIS will be required. Following submittal of the MPO, public scoping will be initiated by the lead federal agency (in this case, BLM) in order to identify potential environmental concerns, or issues. Alternatives to the proposed action will be developed by the BLM, in part from the issues identified during scoping. Given the potentially significant processing times for the NEPA effort, it is advisable to approach the BLM as early in the permitting process as possible.

With regard to Clean Air Act (CAA) permitting, the project area is less than two miles from the Galiuro Wilderness Area, a Class I airshed. CAA permitting would consist of application for a new Title V permit for a new mill; no other aspects of the project are anticipated to require permitting of air emissions. The Title V permit application for this facility would be subject to preconstruction review under the New Source Review provisions, and the permit would have to comply with the Arizona regional haze SIP.

A Clean Water Act (CWA) Section 404 permit is required if a project will result in the discharge of dredged or fill material into jurisdictional waters of the U.S. (waters). The U.S. Army Corps of Engineers (Corps) administers the Section 404 program. The definition of waters is broad; the Corps has commonly asserted jurisdiction over the typically dry arroyos or washes found throughout Arizona. Guidance related to the identification of waters was promulgated in 2007 (and revised in 2008) which has had the effect of limiting the scope of Corps jurisdiction over some ephemeral washes. A jurisdictional waters determination will be completed for the Copper Creek Project area to determine the extent of Corps jurisdiction, though it is anticipated that at the very least Copper Creek itself is likely jurisdictional. Given the proposed impacts to Copper Creek, it is reasonable to assume that an Individual Permit will be required for the project. Significant components of the Individual Permit include the 404(b)(1) alternatives analysis and habitat mitigation and monitoring plan. Once the jurisdictional waters determination has been completed and approved by the Corps, the scope of the Section 404 permitting effort will be determined. Again, it is anticipated that an individual permit will be required.


SGS FORM No. A263a-2/14/13 !

! 9

Section 7 of the Endangered Species Act (ESA) requires that, for any federally permitted action, the permitting authority must evaluate the potential impact of a project to federally listed species, and if an impact is anticipated, consult with the United States Fish and Wildlife Service (USFWS). For the development of the Copper Creek Project, consultation with USFWS would be required for either the CWA Section 404 permit or the Mine Plan of Operations for the BLM. It is anticipated that the BLM would act as the lead federal agency for the purpose of the Section 7 consultation. Given the location of the project area and the nature of the onsite habitat, three federally listed species are anticipated to have a reasonable potential to occur on the project area: southwestern willow flycatcher (Empidonax traillii extimus; SWWF), lesser long-nosed bat (Leptonycteris yerbabuenae; LLNB), and Chiricahua leopard frog (Rana chiricahuensis). In addition, the Sonoran desert tortoise (Gopherus morafkai) is a candidate for listing, which may be listed as soon as 2015.

As with the ESA, federal permitting authorities are required to evaluate the potential for any proposed project to adversely impact properties listed, or eligible for listing, on the National Register of Historic Places (NRHP). A large portion of the project area has been surveyed for cultural resources. A number of the historic mining features within the project area have been identified by archaeological consultants as eligible for listing, either separately or as a single site. Any adverse effects to these features would require consultation and, likely, mitigation.

1.7.2.2 State Permitting

In general, an Aquifer Protect Permit (APP) will be required for any activities or facilities that may result in a discharge of a contaminant to an aquifer, including tailings and waste rock facilities, and non-stormwater impoundments. A review of the APP permitting materials for AMT’s project indicates that AMT was expecting to encounter some acid-generating waste rock that would need to be temporarily stockpiled on the surface. They had agreed to routinely sample, segregate and special-handle this material, to stockpile it on a constructed foundation, and then return it to the excavation site prior to 5 elapsed years. In addition, runoff diversion ditches associated with waste rock stockpiles were required to be lined. The APP process can be time and labor intensive, and for mining projects generally requires extensive hydrologic characterization, engineering (i.e. BADCT demonstration), geochemical characterization, and reclamation design. Completion of the APP process can take up to two years, or more.

In 2011, Redhawk acquired right-of-way access across lands managed by the Arizona State Land Department (ASLD) in support of the development of the exploration decline near a minor tributary of Saloon Gulch. The right-of-way includes the majority of Copper Creek Road as well as the construction of a new all-weather access road that avoids Copper Creek itself. For the larger mine, it is assumed that the access roads will need to be upgraded beyond what is considered in the current ASLD right-of-way in order to accommodate the anticipated increase in the volume and type


SGS FORM No. A263a-2/14/13 !

! 10

of traffic required for the mining effort. In that case, a revision to the right-of-way will be required through coordination with ASLD. As part of the right-of-way application process, a native plant inventory and cultural resources survey will be required.

The State of Arizona requires that mining projects greater than five (5) contiguous acres on private lands submit a reclamation plan to the State Mine Inspectors Office. This plan must outline planned reclamation activities and costs. The reclamation requirements under this program are different than those for the APP program, described above.

The Copper Creek operations are currently covered under the Arizona Pollutant Discharge Elimination System (AZPDES) Multi-Sector General Permit (MSGP-2010) Storm water program administered by the Arizona Department of Environmental Quality (ADEQ), for the historic mining operations within the project area. It is anticipated that the proposed project will not require coverage under an individual AZPDES permit.

Critical path items related to environmental permitting for the project are:

▪ BLM Approvals (including NEPA, ESA, and NHPA compliance), ▪ Clean Air Act permitting ▪ CWA Section 404 permit (including NEPA, ESA, and NHPA compliance), ▪ Aquifer Protection Permit

In general, other approvals are readily acquired within the critical path.

1.8 Capital Cost

1.8.1 Mining

The capital required to develop and sustain a 25,000 ton per day Post-Pillar Cut & Fill mining operation for extracting Copper Creek material from the Childs, Mammoth, Keel, and American Eagle deposits has been estimated to a +/- 35 percent accuracy. These costs are presented in first quarter 2013 US dollars.

The costs are based on unit prices applied to estimated development quantities, recent equipment quotes from mining equipment manufactures, and labor cost build-ups for the mine site location, similar installations, and in-house information on labor productivities for large-tonnage underground room and pillar operations.

The following 1.4 provides a summary breakdown of the mining capital required to develop and sustain a 25,000 tons per day, fully integrated mining and processing facility at Copper Creek, over an average 18 year mine life.


SGS FORM No. A263a-2/14/13 !

! 11

1.8.2 Process Plant and Infrastructure Capital Costs

Process plant and infrastructure capital costs were developed for several throughput levels and a decision was made to fully cost a flotation facility capable of processing 25,000 tons per day. The following Table 1.5 provides a summary of the capital cost items associated with the selected froth flotation facility.

Process plant capital details are specified in the report text and appendices and include details for direct costs, indirect costs, and included a contingency of 35 percent of the sum of the direct plus indirect costs.

Table 1.4 Mine Capital Cost Summary

Cost Center

Preproduction Capital

Estimate ($)

Sustaining Capital

Estimate ($)

Total Capital

Estimate ($)

Mine:

Breccia Decline 5,040,000 0 5,040,000

Shaft Site Prep 3,000,000 0 3,000,000

Prod. & Vent Shafts 163,680,000 0 163,680,000

Level Development 64,283,000 0 64,283,000

Vent/Rock Passes 12,000,000 25,500,000 37,500,000

Internal Ramping 0 40,800,000 40,800,000

Miscellaneous. Small AFE’s 2,000,000 8,500,000 10,500,000

Equipment 118,161,000 73,880,000 191,241,000

Subtotal Mine 368,840,000 148,680,000 517,544,000

Total Estimated Mine Capital 541,918,000 200,967,000 742,885,000

Table 1.5 Summary Process Plant - Capital Cost for Base Case 25,000 tpd

Description Total Cost, US$

DIRECT COSTS

Area 10 – Primary Crushing 10,000,477

Area 15 – Grinding 37,372,924

Area 40 – Copper Moly Flotation 12,174,266

Area 45 – Moly Flotation 3,548,636

Area 50 – Copper Concentrate Handling 2,327,903

Area 55 – Moly Concentrate Handling 3,239,561

Area 60 – Reagents 2,497,033


SGS FORM No. A263a-2/14/13 !

! 12

A total capital cost of approximately US$ 298.7 million was developed for the

Redhawk process plant facilities and infrastructure. The total capital cost includes US$ 177.8 million for total direct costs, US$ 43.4 million for total indirect costs and US$ 77.4 million for contingency.

Area 70 – Tailing Handling 2,005,661

Area 80 – Reclaim and Process Water 1,031,355

Area 90 – Fresh Water 661,312

SUB-TOTAL INSTALLED EQUIPMENT COST 74,859,128

Process Piping 8,020,621

Electrical 13,367,702

Instrumentation 8,020,621

Site Development 10,694,161

Buildings 13,367,702

Roads (Paved, 10.6 miles) 25,440,000

Overhead Transmission Line 16,041,242

Utilities and Outside Lines 8,020,621

SUB-TOTAL DIRECT 177,831,797

INDIRECT COSTS

EPCM 10,694,161

Construction Indirect Costs include: 17,783,180

Construction Supervision Included

Equipment Rental Included

Field Office Expense Included

Mobilization / Demobilization Included

Consumables Included

Owner’s Costs 5,347,081

Spare Parts 2,673,540

Initial Fill & Reagents 2,667,477

Equipment Insurance & Freight Cost 4,277,664

SUB-TOTAL INDIRECT 43,443,103

TOTAL DIRECT AND INDIRECT 221,274,900

Contingency – 35% 77,446,215

TOTAL COST 298,721,115


SGS FORM No. A263a-2/14/13 !

! 13

1.8.3 Tailings Storage Facility (TSF) Costs

Initial and sustaining capital costs for the tailings storage facility are estimated approximately $74.3 million over the life of the facility (without EPCM and contingency). The primary cost is associated with embankment fill. A cost of $4.00 per cubic yard has been assumed due to the nature of the TSF location and the anticipated local availability of fill material. The PEA level cost estimate includes provisions for site preparation, under-drainage collection, a lined external seepage collection pond and reclaim pump works. Costs associated with delivery and water reclaim pipelines between the TSF and the process facilities are included in the process plant cost estimate.

1.8.4 Total Project Capital Cost Summary

The total project capital costs are summarized in Table 1.6 below separated by area of responsibility and separated into pre-production, replacement and sustaining costs.

1.9 Operating Cost

1.9.1 Mining

Mine operating costs have been estimated, based on producing 25,000 tons per day, using a post-pillar cut and fill mining method, from four separate deposits, simultaneously. The estimate was based on prevailing wage rates in the U.S copper mining industry, and underground equipment operating costs in average ground conditions. An electric power cost of US$ 0.077/Kwh was used for estimating power costs. All costs are in first quarter of 2013 US dollars. A summary of this estimate is provided in the following Table 1.7.

Table 1.6 Summary of Total Project Capital Costs (Million US$)

Area Pre-Production Capital

Sustaining Capital/Closure Costs

Mine development cost total 541.9 201.0

Primary 12' Diversion Tunnel 0.0 37.9

Tailing Dam with Contingency 16.4 87.5

Closure Costs 0.0 16.4

Plant and equipment 298.7 0.0

Total Capital Costs 857.1 342.8

Table 1.7 Full Production

Mine Operating Cost Summary


SGS FORM No. A263a-2/14/13 !

! 14

1.9.2 Process

An operating cost estimate for the process plant and associated infrastructure items was developed using criteria similar to those for the mining operation. This estimate is summarized in the following Table 1.8.

Process operating cost estimate for a two-product milling operation at 25,000

tons per day is US$ 4.98/t. Details are given in the text of this report.

1.9.3 General and Administrative

Direct CostsLabor (US$/t)

Supplies (US$/t)

Total (US$/t)

Production 4.15 8.59 12.74

Development 0.32 0.27 0.59

Service & Support 1.30 2.45 3.75

Subtotal Direct 5.77 11.31 17.08

Indirect Costs

Supervision/Technical 0.31 0.06 0.38

Subtotal Indirect 0.31 0.06 0.38

Total Mine 6.09 11.37 17.45

Table 1.8 Summary of Plant Operating Cost by Cost Item

ItemAnnual

Cost (US$)Cost

(US$/ton)

Power 13,231,043 1.47

Labor 7,651,171 0.85

Reagents 4,668,350 0.52

Grinding media 10,539,310 1.17

Repair materials and operating supplies 2,138,832 0.24

Liners and wear materials 5,319,736 0.59

Water supply 1,248,114 0.14

Total 44,796,556 4.98


SGS FORM No. A263a-2/14/13 !

! 15

G&A costs for labor and supplies that will be incurred during the production years and cannot be charged directly to the mine or process were estimated at $0.62 per ton milled.

1.10 Economic Analysis

A total of 155.9 million tons of 0.77 percent copper, 0.015 percent molybdenum, and 0.056 oz silver per ton would be minded over 18 years at the Copper Creek


SGS FORM No. A263a-2/14/13 !

! 1

Project. Pre-production capital cost of US$ 857.1 million and sustaining capital and closure costs of US$ 342.8 million were estimated. The pre-tax economic parameters at various discount rates are summarized in Table 1.9.

Table1.9 above assumes values of US$ 3.00 per lb of copper, US$ 12.00 per lb of molybdenum, and US$ 20.00 per ounce of silver. Economics of the project are highly leveraged to copper price. The economic analysis presented in this PEA is based on mineral resources. This PEA is preliminary in nature and includes inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves. There is no certainty that the preliminary economic assessment will be realized. Mineral resources that are not mineral reserves do not have demonstrated economic viability.

Additional details concerning the pre-tax cash flow analysis are found in Section 22.

1.11 Conclusions and Opportunities

The following measures are recommended in order to improve the economic viability of the project:

Table 1.9 Summary of Key Financial Parameters

Item Pre-Tax

Payback, operating years 6.2

Net Present Value (NPV) (million US$)

0.00% $1,311

5.00% $457

7.50% $231

10.00% $79

Internal Rate of Return (IRR) 11.8%


SGS FORM No. A263a-2/14/13 !

! 2

1.11.1 Mining

Additional exploration drilling to fill in gaps in the existing resource;

▪ Create larger, more consistent, above cutoff grade mining areas ▪ Increase current resource Measured and Indicated tons ▪ Convert current Inferred tons to Measured and Indicated tons ▪ Potential to improve grades ▪ Potential to find additional deposits and expand existing deposits

Additional exploration in the many unexplored breccias pipes that exist in the area;


SGS FORM No. A263a-2/14/13 !

! 1

▪ Additional tons at higher grades ▪ Potential for new mining areas, at shallower depths

Perform additional deep exploration;

▪ Potential for large porphyry deposit at depth

11.11.2 Milling

Further Metallurgical Testing and studies are recommended to improve;

▪ Plant Design ▪ Recovery ▪ Concentrate Grade ▪ Operating Costs - additional studies are required to attempt to reduce power

consumption since this will be a significant operating cost.

1.11.3 Infrastructure

▪ Power: The option for on-site power generation was considered; however, the cost associated with the supply and transportation of natural gas to the mine site and installing a gas turbine engine power generation system on site requires a detailed study.

▪ Water Supply: The investigation of a source for mine water is in progress. The well field target has been selected in the alluvial pediment of Galiuro Mountains to the west of the mining area on state owned land. The proposed well field is located in the vicinity of an existing state-owned well where preliminary pumping tests yielded encouraging results.

▪ Access Road: Currently there is an unpaved dirt road access to the proposed mine and mill site. For this study it was assumed that this 10 mile access road will be upgraded and paved to improve access to the plant site.

1.11.4 Tailings Storage Facility

There is potential to mine the upper portion of the Old Reliable by open pit methods. The use of mine waste rock for TSF embankment fill could potentially reduce fill costs and decrease the disturbance area associated with fill borrow and waste rock disposal.

The evaluation of future tailings properties could demonstrate the feasibility of alternative construction methods for TSF embankment raises. Upstream or centerline dam raises could reduce embankment fill requirements in later TSF construction phases. Alternative construction methods will have no impact on initial capital costs,


SGS FORM No. A263a-2/14/13 !

! 2

however; sustaining capital costs could be significant reduced if fill requirements are reduced.


SGS FORM No. A263a-2/14/13 !

! 1

2.0 INTRODUCTION

Redhawk Copper, Inc. (Redhawk) has commissioned SGS Metcon/KD Engineering (SGS) of Tucson, Arizona to provide a scoping level study for the Copper Creek Project located 75 road miles northeast of Tucson, Arizona.

Contributions to this report were made by a number of organizations and individuals.

Mr. Joseph M. Keane, P.E. has overall responsibility for the report contents and specifically parts 1.5, 1.7, 1.8, 1.9, 1.10 and 1.11; Section 2 and Section 3; Section 13 & 17, 18.1, 18.2, 19, 20.3, 21.3; Sections 22 and 23, as well as Sections 24 through 27. Mr. Keane has not visited the property.

Mr. Herb Welhener of IMC in Tucson, AZ is responsible for parts of 1.1, and 1.2; Sections 4 through 12, and 14 and 16.2. Geological contributions to the development study were made by Mr. Welhener. Mr. Welhener visited the property on 4 October 2006, 30 October 2009 and again on 10 September 2011.

Mr. Steve Milne, P.E. of Milne & Associates, Inc. (MAI) is responsible for Section 1.3 1.4 and 1.4.2; Section 15, 16.1, 16.4 through 16.6; 21.1 and 21.1.1, of the report. Mr. Milne visited the property on 18 November 2005.

Mr. Gene Muller, P.E. of Golder Associates (GAI) is responsible for the preparation of Sections 1.6; Sections 18.3 through 18.7; 20.1, 20.2, 21.1.2 and 21.1.3, 21.4 through 21.6. Mr. Muller visited the property 18 July 2006 and again 27 February 2009.

Mr. David Nicholas of Call and Nicholas, Inc. (CNI) is responsible for parts 1.4.1 and 16.3..

Additional important contributions to the report were provided by SGS personnel and those employed by Redhawk.

Table 2.1 below lists the responsibilities for the Qualified Persons (QP) as defined by the NI 43-101 Requirements.


SGS FORM No. A263a-2/14/13 !

! 2

Table 2.1 Summary of Qualified Persons

Report SectionCompany/

Author QP

1 Summary

1.1 Property Description and Location IMC Herb Welhener

1.2 Geological Setting and Deposit Type IMC Herb Welhener

1.3 Mineral Resource Estimates MAI Steve Milne

1.4 Mining MAI Steve Milne

1.4.1 Geotechnical Evaluation CNI David Nicholas

1.4.2 Basic Mine Plan MAI Steve Milne

1.5 Mineral Processing and Metallurgical Testing SGS J. M. Keane

1.6 Tailings Disposal Facility GAI Gene Muller

1.7 Environmental Studies, Permitting and Social or Community Impact SGS J. M. Keane

1.8 Capital Costs SGS J. M. Keane

1.9 Operating Costs SGS J. M. Keane

1.10 Economic Analysis SGS J. M. Keane

1.11 Conclusions and Opportunities SGS J. M. Keane

2 Introduction SGS J. M. Keane

3 Reliance on Other Experts SGS J. M. Keane

4 Property Description and Location IMC Herb Welhener

5Accessibility, Climate, Local Resources, Infrastructure and Physiography IMC Herb Welhener

6 History IMC Herb Welhener

7 Geological Setting IMC Herb Welhener

8 Deposit Types and Mineralization IMC Herb Welhener

9 Exploration IMC Herb Welhener

10 Drilling IMC Herb Welhener

11 Sample Preparation, Analysis and Security IMC Herb Welhener

12 Data Verification IMC Herb Welhener

13 Mineral Processing and Metallurgical Testing SGS J. M. Keane

14 Mineral Resource Estimates IMC Herb Welhener

15 Mineral Reserve MAI Steve Milne


SGS FORM No. A263a-2/14/13 !

! 3

16 Mining

16.1 Tonnage Estimate for Post Pillar Cut and Fill Mining MAI Steve Milne

16.2 Cut Off Evaluation IMC Herb Welhener

16.3 Geotechnical Considerations CNI David Nicholas

16.4 Basic Mine Plan MAI Steve Milne

16.5 Preproduction Infrastructure / Development / Equipment MAI Steve Milne

16.6 Production MAI Steve Milne

17 Recovery Methods SGS J. M. Keane

18 Infrastructure

18.1 Access Roads/Project Roads SGS J. M. Keane

18.2 Power Supply and Distribution SGS J. M. Keane

18.3 Water Supply and Distribution GAI Gene Muller

18.4 Project Water Balance GAI Gene Muller

18.5 Tailings Impoundments GAI Gene Muller

18.6 Diversion Tunnel GAI Gene Muller

18.7 Hydraulic Backfill Plant GAI Gene Muller

19 Market Studies and Contracts SGS J. M. Keane

20Environmental Studies, Permitting & Social or Community Impact

20.1 Hydrogeological GAI Gene Muller

20.2 Geochemical Considerations GAI Gene Muller

20.3 Environmental Permitting SGS J. M. Keane

21 Capital and Operating Costs

21.1 Underground Mine MAI Steve Milne

21.1.1 Underground Mine Capital Estimate MAI Steve Milne

21.1.2 Hydraulic Backfill Cost Estimate GAI Gene Muller

21.1.3 Discussion GAI Gene Muller

21.2 Underground Mine and G&A Operating Estimates MAI Steve Milne

21.3 Process Plant SGS J. M. Keane

21.4 Tailings Storage Facility Capital Costs GAI Gene Muller



SGS FORM No. A263a-2/14/13 !

! 4

21.5 Diversion Tunnel Costs GAI Gene Muller

21.6 Reclamation and Closure GAI Gene Muller

22 Economic Analysis SGS J. M. Keane

23 Adjacent Properties SGS J. M. Keane

24 Other Relevant Data and Information SGS J. M. Keane

25 Interpretation and Conclusions SGS J. M. Keane

26 Recommendations SGS J. M. Keane

27 ReferencesAs

requiredAs

required



SGS FORM No. A263a-2/14/13 !

! 1

3.0 RELIANCE ON OTHER EXPERTS

Subject to normal due diligence, SGS has relied on the accuracy of reports and data supplied by Redhawk and other geological and mineral engineering consultants in the preparation of the Independent Report. SGS has reviewed and analyzed data provided by Redhawk and other geological and mineral engineering consultants, and has drawn its own conclusions there-from, augmented by its direct field examinations. SGS has not carried out any independent exploration work, drilled any holes or carried out sampling or assaying on the property.

The authors acknowledge the full cooperation of management and field staff from Redhawk, all of whom made any and all data requested available and responded openly and helpfully to all questions, queries and requests for material. All maps, as well as certain Tables and Figures for this report were supplied by Redhawk.


SGS FORM No. A263a-2/14/13 !

! 2

4.0 PROPERTY DESCRIPTION AND LOCATION The Copper Creek property is located in Pinal County, Arizona on the east flank of the Galiuro Mountains. The property is approximately 75 miles northeast of Tucson, Arizona and 15 miles from San Manuel, Arizona. The property ownership is comprised of approximately 1150.161 acres of Patented Federal Mining claims, 12 Arizona State Prospecting Permits totaling 4,490.61 acres, 754 unpatented Federal Lode Mining claims owned by Redhawk and four unpatented Federal Lode Mining claims under lease purchase agreement to Redhawk totaling 14,456.83 acres, and 779 acres of private lands. The total acreage controlled by Redhawk on the Copper Creek property is more than 20,713 acres. The property is situated in the Bunker Hill Mining District.

4.1 Location

The approximately 20,713 acre Copper Creek Property is located about 75 road miles northeast of Tucson. The Property lies approximately 10 miles from the town of Mammoth, in the Bunker Hill Mining District on the western slopes of the Galiuro Mountains, Pinal County, Arizona, USA (Figure 4.1). It lies in sections 1, 2, 10, 11, 12, 13, 14, 23, 24, T7S R17E, sections 17, 18, 19, 20, 21, 27, 28, 29, 30, 31, 32, 33, 34, and 35 T7S R18E, sections 1, 2, 3, 4, 9, 10, 11, 12, 13, 14, and 15 T8S R18E G&SRB&M and is centered at 32° 45' N Latitude, 110° 30' W Longitude.

Copper Creek is situated somewhat central to the copper mining districts in Arizona with the Globe - Miami district to the north, the Mission-Pima-Sierrita district to the south and the Safford and Morenci districts to the east (Figure 4.2).


SGS FORM No. A263a-2/14/13 !

! 3

!

Figure 4.1 - Location Map


SGS FORM No. A263a-2/14/13 !

! 4

! Figure 4.2 - Arizona Copper Infrastructure

4.2 Land and Tenure and Property Agreements

The Copper Creek Property comprises one contiguous group of 61 patented and 758 unpatented Federal claims, twelve Arizona State Mineral Exploration Leases, and private lands. All the above mineral titles are held directly by Redhawk Copper, (USA) Inc. (Figure 4.1). The claims information in Appendix 1 is not a legal title opinion but is a compilation of claims data based on information supplied to the author by Redhawk Copper, (USA) Inc. Redhawk has informed the author that the claims and prospecting permits are properly maintained and all appropriate fees have been paid. Based on this


SGS FORM No. A263a-2/14/13 !

! 5

information, the claims and leases appear to be in good standing as of the date of this report.

To renew unpatented claims, an Annual Maintenance Fee of US$ 140 per claim levied by the Federal Bureau of Land Management (BLM) must be paid on or before 1 September each year. The current annual maintenance cost of the unpatented claims is approximately US$ 106,120. In addition, there is a small yearly County filing fee for the unpatented claims. There are currently twelve Arizona State mineral exploration permits that belong to the Copper Creek landholdings. These permits are valid for five years, assuming the work requirements are fulfilled each year. All twelve leases, which expire mid-2016, are in the name of Redhawk Copper, Inc. There is a renewal fee of US$ 100 each per year. In addition, there is a fieldwork requirement due (or cash in lieu of work) for each lease, and that is in the amount of US$ 10 per acre for years 2 and 3 and US$ 20 per acre for years 4 and 5. The fifteen permits encompass a total of 4,490.61 acres. The work must be done on these permits, or approximately US$ 44,491 must be paid as cash in lieu. The patented claims and some of the unpatented claims have been surveyed; the permits have not been surveyed (their location is determined by a legal description (i.e. Section, Range, Township, etc). Annual property taxes to Pinal County are currently US$ 9,667. Surface title over the area of the Federal unpatented claims is comprised of BLM, State of Arizona Lands, and private ownership under the Stock Raising Homestead Act. State exploration permits are under State of Arizona jurisdiction. Unpatented mining claims are under BLM jurisdiction. The surface title over the patented claims is held by Redhawk Copper, Inc. The surface access to carry out work to the BLM and State land is through normal BLM and State permitting procedures respectively.


SGS FORM No. A263a-2/14/13 !

! 6

5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

The Copper Creek Property is located in a ranching and mining area about 75 road miles northeast of Tucson, Arizona. Tucson is a major population center and transportation hub with well-developed infrastructure and services to support the area mines. Access to the property from Tucson is by paved highway 65 miles to the property road junction, then 10 miles on gravel road to the site. The driving time from Tucson to the site is approximately two hours. A network of ranch and mine roads provides access to most parts of the property.

The Copper Creek Property is located in rugged terrain of the eastern Basin and Range physiographic province of South Eastern Arizona. Elevations on the property range from 3,400 ft to 4,900 ft above mean sea level.

Climate varies with elevation, but over most of the property summers are hot and dry and winters are mild. Data collected over a fifty-year period at a weather station approximately 10 miles from site at an elevation of 3,500 ft indicates the average summer temperatures range from a low of 67.4oF at night to a high of 96oF during the day. During winter, average temperatures range from a low of 35.7oF at night to a high of 61.8oF during the day. The lowest temperature recorded was 12oF and the highest recorded temperature was 111oF. Precipitation data collected from the same weather site indicates the average annual rainfall ranges from 6.3 to 26.8 inches and averages 13.8 inches per year. A rain gauge maintained on Copper Creek since 1987 recorded average annual rainfall of 13.0 inches in a range of 8.94 inches to 24.7 inches. Occasional light snow falls at higher elevations in the winter months. The average annual evaporation rate, approximately 90 inches per year, greatly exceeds the average annual precipitation rate.

Vegetation, including various cactus, mesquite and palo verde trees, grasses and scrub brush, are common in the area.

The area has an exceptionally well-developed mining infrastructure. BHP’s San Manuel operation was located approximately 10 miles to the west of Copper Creek. This operation was shutdown in 2001, but substantial infrastructure including roads, power, and towns are available near Copper Creek. The large operating mine-concentrator-smelter complex owned by Asarco (Grupo Mexico SA) is located at Hayden, 30 miles to the northwest of Copper Creek. Other similar facilities are located throughout Arizona.

The previous property operator carried out ground water studies and determined that adequate water for exploration drilling and any possible future mining operation may be obtained from wells on or near the property.


SGS FORM No. A263a-2/14/13 !

! 7

6.0 HISTORY

6.1 The Early Years

According to old reports, some rich lead-silver ore was mined from the Bluebird vein as early as 1863; however, the Bunker Hill (Copper Creek) mining district was not organized until 1883, after completion of the transcontinental Southern Pacific railroad (1880). Some ore was shipped from the Bluebird vein during the 1880s silver boom, Claims were staked to cover copper deposits prior to 1900, but little work was done in the district until after 1902. Principal freight access was over the crest of the Galiuro Mountains, to Wilcox.

In 1903, Copper Creek Mining Company acquired claims covering narrow copper-silver veins in Tertiary-age volcanic rocks near the headwaters of Copper Creek and later acquired ground along Copper Creek, as far west as the Laramide age Old Reliable pipe deposit. The wagon road from Mammoth to Copper Creek reportedly was constructed in 1908; this provided more favorable freight haul after the railroad extended to Winkleman in 1911. By 1913, the Copper Creek Mining Company and its successors (the Minnesota-Arizona Mining Company and Copper State Metals Mining Company) had constructed a dam, power plant, dispensary and 200 ton per day gravity concentrator in the vicinity of Post Office Point. They had developed and mined a small, tabular breccia ore body at American Eagle; the Old Reliable pipe deposit was partially developed for mining and was connected to the concentrator with about two miles of narrow gauge railroad. By the end of 1913, the company was in default and employees were working the Old Reliable to recover unpaid wages. The company was refinanced in 1914 and operated at Old Reliable where about 30,000 tons were produced prior to shutdown in 1919. The Old Reliable claim group was surveyed for patent in 1919.

Commencing in 1907, the Calumet and Arizona Mining Company (C & A), guided by Ira Joralemon, explored the Copper Giant, Copper Prince, Glory Hole (Globe), and Superior pipes by adits, shafts and drifts. To supplement the underground exploration, C & A drilled about 6,000 ft in fourteen surface holes, during 1914. A copper resource was found in both the Glory Hole and Prince pipes, but there was no production. The C & A group of 26 claims was surveyed for patent between 1908 and 1919; title passed to Phelps Dodge Corporation in 1931 when it purchased the Calumet and Arizona company. The only recorded production from the C & A ground has been by Arizona Molybdenum Corporation which mined and milled 23,312 tons from the Copper Prince pipe with average grade of 3.19 percent copper, during 1937. Written logs of the C & A drill holes reside in Redhawk’s files, but core has not been located.

C & A holes are numbered DH-1 through DH-14. An adit, driven below the outcrops of the Childs Aldwinkle pipes in 1915, discovered the copper-molybdenum ore body there and the pipes were partly developed for production during 1917-1918; the claims were surveyed in 1916 and patented in 1919. Arizona Molybdenum Corporation acquired the property in 1933 and proceeded to


SGS FORM No. A263a-2/14/13 !

! 8

develop the copper and molybdenum ore body to 520 ft below the haulage level. The old Arizona-Minnesota Mining Company gravity concentrator was converted to flotation; about 350 tons per day were processed. In 1935, a new flotation concentrator was constructed on the Childs Aldwinkle property, near the portal of the haulage adit; about 300 tons per day were processed there. Between 1933 and 1938, about 329,000 tons were milled. Leasers worked the mine in 1939 and again during 1957-1965. At some time prior to 1957, the Childs Aldwinkle winze was extended to 680 ft below the haulage adit; the 680 level was developed and six short holes were drilled there by Inspiration Consolidated Copper Company; Magma Copper Company obtained logs of these holes from Inspiration in 1967. These logs are in Redhawk's files, but the core has not been located.

Holes drilled by Inspiration at Childs Aldwinkle: H-1, 2, 3, 4, 5, 6

During his work at Copper Creek for C & A, Ira Joralemon postulated that the area between the chalcocite-enriched Old Reliable and Glory Hole pipes may be underlain by a chalcocite blanket of commercial tenor. In the late 1940's, Copper Creek Consolidated Mining Company (Morris Elsing) secured the patented claims at Old Reliable and, during 1950, drilled four holes with a churn-drill to disprove Joralemon’s idea. Copper Creek consolidated held the Old Reliable property, without recorded production, until about 1954.

Holes drilled by Morris Elsing near Old Reliable: CDH-1, 2, 3, 4

6.2 The Modern Era

In 1956, Siskon Corporation acquired ground that had been part of the more westerly properties of the old Copper State Metals Mining Company, most in Sections 10, 11 and 14. Siskon’s principal interest was the Old Reliable mine. The Old Reliable had been rehabilitated and sampled in 1942 and 1943 by the U.S. Bureau of Mines and very encouraging results reported in RI 4006 (1947). Siskon drilled 21 diamond core holes from the 100 and 200 levels of the Old Reliable mine. Neither logs nor core are available.

Holes drilled by Siskon at Old Reliable: OR1, 2, 3, 4, 5, 6, 7, 8, 9, 10, 11, 12, 13, 14, 15, 16, 17, 18; SW-1, 2, 3

In 1959, Bear Creek Mining Company (Kennecott) optioned the Siskon ground and also the Childs Aldwinkle patented claims. Bear Creek mounted the first integrated exploration (geologic mapping, geochemical and geophysical surveys, followed by drilling), at Copper Creek. Fifteen holes were drilled there by Bear Creek and several of these cut mineralized zones. However, none of the Bear Creek intersections appeared to be minable and they abandoned the project in 1962. The Bear Creek drill core was stored in a cabin at the Old Reliable under care of watchman Pete Carey. This core was moved to a warehouse at Magma’s plant in San Manuel in 1967 and is now in the possession of Redhawk.


SGS FORM No. A263a-2/14/13 !

! 9

Holes drilled by Bear Creek: Old Reliable CU-3, 10 Childs Aldwinkle CA-1, 2, 3

Siskon claims CU-1, 2, 4, 5, 6, 7, 8, 9, 11, 12

In 1966, Newmont Exploration Limited (NEL) optioned the Siskon property and enlisted Magma Copper Company (80.3 percent owned by Newmont Mining Corporation) as co-venture and operator. The Childs Aldwinkle patented claims were also optioned, as were adjacent claims owned by Clark, Downey and Lehman as well as the patented Redbird claims (Bluebird Mine). Additional claims were located to cover open Federal land and State land was leased. Exploration was directed toward discovery of a major disseminated copper deposit; breccia pipes were not a primary target. Between 1966 and 1970, geology of the district was mapped and 30 deep core holes were drilled. Core from these and all other JV holes drilled at Copper Creek was stored in a warehouse at Magma’s plant at San Manuel, until June, 2005. This work demonstrated a significant copper-mineralized zone, at depth, beneath the American Eagle Area. Magma became a wholly-owned subsidiary of Newmont in 1969.

Holes drilled by Magma - Newmont: Siskon claims SK-1, 2, 3, 4, 6, 7 Bonbright claims B-20, 24, 29, 30 (Childs Aldwinkle) Downey claims D-5, 8, 9, 13, 26 Lehman claims L-10, 12, 19, 25, 27 Magma - Newmont claims M-22, 28 A-11 S-14, 15, 16, 17, 18, 21, 23

Although Newmont acquired the Siskon ground in 1966, Siskon retained the right to deal separately with the upper part of the Old Reliable pipe and certain adjacent claims, for a period of 15 years. In 1968, Occidental Minerals Corporation (Oxymin) leased that ground from Siskon and also optioned part of the adjacent Phelps Dodge ground that covered the mineralized Glory Hole, Copper Prince and Copper Giant pipes. The old workings that C & A had driven to test the (Phelps Dodge) pipes were rehabilitated above the water table and Oxymin drilled 67 surface and underground holes to test the Old Reliable, Glory Hole, Prince and Giant pipes. Oxymin released their option on the Phelps Dodge ground in about 1970. Redhawk has copies of Oxymin core logs, but no logs for percussion holes; location of the core is not known.

Holes drilled by Oxymin: Old Reliable OOR1, 2, 3, 4, 5, 6 (Surface – Core) UG-1, 2, 3, 4, 5, 6, 7, 8 (100 level – Core)

EH-1, 2, 3, 4, 5, 6, 7, 8, 9, 10, 11, 12, 13, 14, 15, 16, 17, 18, 19, 20, 21, 22, 23, 24, 25, 26

(100&200 levels-Percussion)

Glory Hole GH-1, 2, 3 (Surface – Core)


SGS FORM No. A263a-2/14/13 !

! 10

EHGH-1, 2, 3, 4, 5, 6, 7, 8, 9, 10, 11, 12, 13, 14, 15 (Underground–Percussion)

Prince OP-1 (Surface – Core) EHOP-1, 2, 3, 4, 5, 6 (Underground-Percussion)

Giant OG-1, 2 (Surface – Core)

Oxymin, in 1971, assigned their interest in the Old Reliable to Ranchers Exploration and Mining Company. Ranchers drilled three holes to confirm results of Oxymin drill holes; logs of the Ranchers holes are in Redhawk’s file, but location of the core is not known. In 1972, Ranchers, in association with Du Pont, rubblized the Old Reliable pipe above the 3730 elevation, by blasting with ANFO. Copper was leached from this rubble-column with dilute sulfuric acid; copper was recovered from the leach-liquors by precipitation on tin-cans in a plant below the mine. More than 12,077,000 pounds of cement copper were recovered between 1972 and 1981, when Ranchers’ lease expired.

Holes Drilled by Ranchers at Old Reliable: RD-1, 2, 3

In the years 1972 - 1974, after retrieving its property from Oxymin, Phelps Dodge (now Freeport Copper & Gold, Inc.) geologists mapped, sampled and tested the Phelps Dodge ground with geophysics; nine holes were drilled to test deep targets, with disappointing results. Phelps Dodge did not explore the pipe deposits on its ground. Redhawk has copies of the Phelps Dodge drill hole logs. Phelps Dodge has the core.

Holes drilled by Phelps Dodge on their ground: CC-1, 2, 3, 4, 5, 6, 7, 8, 9

Humble Oil joined Newmont and Magma in exploration for porphyry copper deposits at Copper Creek in 1971. Humble assumed project management during 1971-1972, their “earn-in” period and drilled 20 deep holes. It was Humble-Newmont hole HN-12 that discovered the third (north) finger of the Childs Aldwinkle pipe. Redhawk has both logs and core.

Holes drilled by Humble – Newmont: American Eagle HN-1, 2, 3, 4, 5, 6, 7, 8, 9, 10, 11, 16, 19, 20 Childs Aldwinkle HN-12, 14, 17, 18 Old Reliable HN-13 Joint Venture claims HN-15

Humble Oil was renamed Exxon Corporation in about 1973. In 1979, Exxon, at their sole cost, drilled a hole to test a geological theory that was advanced by their exploration management (but failed to find ore). Redhawk has both log and core.

Hole drilled by Exxon (JV ground): EN-1


SGS FORM No. A263a-2/14/13 !

! 11

Newmont resumed management of the Copper Creek Joint Venture in 1973 and drilled six angled holes from surface to test pipe targets. However, by the mid-1970s, Newmont’s corporate interest in porphyry copper exploration had waned. The Copper Creek project reverted to care and maintenance; drill targets were carefully selected, but drilling was reduced to the amount needed to underwrite property maintenance costs. Hole NE-6 discovered the lower Mammoth feeder-zone and hole NE-10 discovered the Mammoth breccia pipe. Redhawk has both logs and core for these holes. Between 1972 and 1977, the joint venture surveyed for patent claims on public domain that would be interior to a crack-line projected at 45 degrees from the bottom of the American Eagle deposit. This survey was filed with the BLM but the claims were not patented. Exxon ceased contributing to the joint venture in 1985 and withdrew in 1987.

Holes drilled by Newmont-Exxon: Childs Aldwinkle AH-1, 2, 3 American Eagle AH-4, 5, 6 NE-2, 4, 7, 8 Mammoth area NE-5, 6, 10 Railroad pipe NE-9 Joint Venture claims NE-1, 3

When Newmont distributed Magma’s equity to Newmont’s shareholders in 1987, Newmont’s ownership interest in properties at Copper Creek was incorporated into Magma and Magma became an independent company. Magma’s management had little interest in exploration at Copper Creek; they reduced the size of the property package, but held the core property. Magma met the requirement for assessment expenditure by drilling three holes. Redhawk has logs and drill core for these holes.

Holes drilled by Magma: Mammoth area CC-1 Childs Aldwinkle CC-2, 3

6.3 Work At Copper Creek By AMT (USA) Inc.

Arizona Mineral Technology (Kushal Singh) finalized an agreement to acquire the Copper Creek property from Magma in 1994. Singh’s company was renamed AMT International Mining Corporation when it was incorporated in Canada.

Between 1960 and 1995 when AMT became active at Copper Creek, more than 77 deep holes had been drilled there by major copper companies and large amounts of geological, geophysical, geochemical and other analytical data had been generated by them. These explorers were searching for a major porphyry copper deposit; the mineralized pipes were too small to interest these major companies.

AMT began field investigations at Copper Creek in the spring of 1995; the relatively shallow mineralized pipes were favored targets for AMT. Claims were staked to recover the ground dropped by Magma and to fill fractions. Agreements to acquire the Bell (Ryland) ranch and the Mercer ranch were signed. An agreement was signed with


SGS FORM No. A263a-2/14/13 !

! 12

Phelps Dodge Exploration Corporation to obtain an interest in their patented claim block. AMT obtained a prospecting permit for State lands in the south half of Section 2, near the Bluebird mine. Access and drill roads were repaired and a new access road was constructed from Saloon Gulch to the top of White Bear hill.

During May and June 1995, AMT drilled nine reverse circulation (RC) holes at Old Reliable to confirm that leaching by Ranchers had not significantly depleted the chalcocite ore body there. In addition, three RC holes were drilled at Old Reliable in June 1996 and 20 RC holes were drilled there in January through March, 1997. Six of these vertical RC holes were extended with the core drill to test the deposit below the rubble column. These holes plus pre-AMT holes and rock sample assays comprise the data-base from which the resource in the Old Reliable pipe has been estimated.

Holes drilled by AMT at Old Reliable: OR-1R (core), 2R, 3R, 4R, 5R, 6R, 7R, 8R, 9R, 10R, 11R, 12R,

13R (core), 14R (core), 15R (core), 16R (core), 17R (core), 18R, 19R, 20R, 21R, 22R, 23R, 24R, 25R, 26R, 27R, 28R, 29R, 30R, 31R, 32R

AMT drilled 40,135 ft in 37 angled diamond core holes to test the Childs Aldwinkle pipe above 2800 elevation, in March through September, 1996. These westerly-directed holes were drilled from four surface sites, east of the Childs Aldwinkle glory-holes. In addition, 3580 ft were drilled in nine RC holes to test the top of the blind north finger of the pipe and three vertical core holes were drilled to obtain metallurgical test samples. These holes, plus pre-AMT holes, comprise the data-base from which the resource in the Childs Aldwinkle pipe deposit has been estimated.

Holes drilled by AMT at Childs Aldwinkle: CA28+3A, CA28+4, CA28+5, CA28+8, CA30+3, CA30+4, CA30+5, CA30+6, CA32+3, CA32+4, CA32+5, CA32+6, CA32+8, CA34+2, CA34+3, CA34+4, CA34+4A, CA34+7, CA35.5+1, CA35.5+2, CA35.5+3, CA35.5+4, CA36+7, CA36+8, CA36.5+2, C36.5+3, CA36.5+4, CA37.5+1, CA37.5+1A, CA37.5+2, CA37.5+3, CA38+6, CA38+7, CA40+6, CA40+7, CA40+8, CATECH, Met-2CA, 3CA, 4CA

CA-1R, 2R, 3R, 4R, 5R, 6R, 7R, 8R, 9R

In the years 1976 through 1982, Newmont drilled two core holes beneath the south wall of Copper Creek canyon, almost directly beneath the prior location of the Arizona Molybdenum Corporation concentrator. Both of these holes cut mineralized intervals, of potential ore-grade. The near surface intercept in hole NE-10 was similar to other copper-mineralized breccia-pipe deposits in the area, but the deeper intercepts in hole NE-6 were pervasive sericite-chalcopyrite replacement of granodiorite. Follow-up holes drilled by AMT, 37,578 ft in 24 angled and seven vertical holes, defined the Mammoth breccia, a pipe-form, quartz-chalcopyrite veinlet stockwork deposit with N25W elongation; potentially economic parts of the pipe, bottom above the 2800


SGS FORM No. A263a-2/14/13 !

! 13

elevation. These 31 core holes drilled by AMT, plus Newmont’s hole NE-10, comprise the data-base from which the Mammoth pipe resource was estimated.

Holes drilled by AMT to test the Mammoth pipe: CK32+0, 33+1Y, 33+3Y, 34.5-50, 35+3Y, 35.5-50, 36+0, 37+100, 37+50 37.2+50 CK-B, C, D, E VIX24-2, 28-1, 28-2, 30-1, 32+1, 32-1, 32-2, 32-3, 34+1, 34-1, 34-2, 36-2 VIX-A, VIXTECH, UM-1, MET1-CK, MET5-CK

During November, 1996, hole VIX 28-2 extended through the relatively shallow Mammoth pipe and into sericite-chalcopyrite rock, similar to the deep mineral intercept in Newmont hole NE-6. The similarity between these mineral intercepts, about 700 ft apart, stimulated drilling of holes to test the continuity of the intervening “Lower Mammoth” (LM) mineralized zone. The Lower Mammoth deposit is a steep, N25W- trending, altered and mineralized sheared zone that clearly has fed mineralizing fluids upward, into the Mammoth pipe. In addition to Newmont’s hole NE-6, the Lower Mammoth zone has been tested by an additional 13 holes drilled by AMT; it is, however, incompletely drilled at this juncture. The limited current information indicates potentially economic parts of the Lower Mammoth feeder zone tops near 2400 elevation; it appears to have reasonable continuity for at least 700 ft along strike, to be open downward and to the southeast. Both tenor and thickness appear to increase toward the south, where thickness of the zone probably exceeds 100 ft.

Holes drilled by AMT to test the Lower Mammoth zone: VIX24-2, 28-2 LM-1, 2, 3, 4, 5, 6, 7, 8, 9, 10, 11

In addition to holes drilled to test the four major ore bodies described previously, AMT drilled a number of core and RC holes to test other mineral occurrences at Copper Creek. Some of these holes returned interesting (but generally sub-economic) assays that warrant follow-up. Hole logs, split drill core and RC cuttings trays are in possession of Redhawk.

Holes drilled by AMT to test various targets at Copper Creek (partial list): American Eagle AE-1, 2, 3 AE-1R, 2R, 3R B-24 Breccia B24-1R, 2R, 3R Boomerang breccia BG-1R Bluebird area S2-98-1 Copper Giant CG1, 2, 3, 4, 5, 6

Copper Prince CP1, 2, 3, 4, 5, 6 Doreen breccia DB1, 2, 3, 4 Glory Hole (Globe) G1, 2, 3, 4, 5, 6, 7, 8, 9, 9A, 10, 11, 12, 13 M22 drill site M22-1R Marsha breccia MB-1, 2, 2A Pole breccia PB30+14, 34+14, 38+14


SGS FORM No. A263a-2/14/13 !

! 14

PB-1R, 2R, 3R, 4R, 5R Railroad pipe RR-1R, 2R, 3R, 4R, 5R PC1

Post Office breccia PO-1R Rum claims RUM-1 Shirley breccia SB-1, 2 Superior breccia S1, 2, 3 White Bear pipe WB-1R, 2R, 3R, 4R, 5R

AMT mounted a staggered, 200 ft grid, which covered much of the productive ground at Copper Creek. This grid was used to guide collection of geochemical samples, ground magnetic and radiometric surveys. The exploration survey data is in files controlled by Redhawk and warrants study.

AMT exhausted its financial resources in 2001 and ceased all exploration. Norshield Investments, AMT’s primary creditor, advanced funds necessary to maintain the key properties at Copper Creek, but agreements to secure the “Ryland” ranch, the Mercer Ranch, the Phelps Dodge claims and Downey’s Moose claims were dropped.

Redhawk Resources reviewed the project data in 2004. Redhawk acquired AMT’s remaining property at Copper Creek, as well as the drill core, rock samples and the accumulated project data, in 2005.

4. Redhawk Resources

Redhawk Resources reviewed the project data in 2004. Redhawk acquired AMT’s property at Copper Creek, as well as the drill core, rock samples and the accumulated project data, in 2005. Redhawk spent considerable time following the acquisition organizing and consolidating the available data and drill core. The core is now housed in a core storage facility at Redhawk's project office in San Manuel, Arizona. During the review of the drill core, it was discovered that some of the last holes drilled by AMT were not split or assayed. Redhawk has logged, split and assayed this core and the results were added to the data base.

Redhawk commissioned Independent Mining Consultants, Inc. (IMC) to develop a resource estimate of four of the mineralized targets (Mammoth Breccia, Childs-Aldwinkle Breccia, Old Reliable Breccia and the Lower Mammoth – Keel deposit) on the Copper Creek ground. IMC’s work started in March 2006 and a resource was announced in September 2006. The NI 43-101 Technical Document for this resource is titled “Copper Creek Property Mineral Resource, Pinal County, Arizona, USA, Technical Document”, dated October 31, 2006 and filed on SEDAR on November 8, 2006.

Redhawk commenced a drilling program in 2006 in the Mammoth Breccia deposit and in the breccia pipes located on the claims acquired from Phelps Dodge Corporation.

In 2007, Redhawk commissioned IMC to develop a resource estimate for the


SGS FORM No. A263a-2/14/13 !

! 15

American Eagle deposit to report the resource and to provide guidance to Redhawk’s drilling in the American Eagle. The resources are documented in the IMC’s NI 43-101 Technical Report titled “American Eagle Deposit Mineral Resource, Copper Creek Property, Pinal Country, Arizona, USA, Technical Report” dated November 26, 2007 and filed on SEDAR on November 29, 2007.

In late 2007 and early 2008 Redhawk completed twelve rotary hammer pre-collar holes totaling 18,024 feet in the American Eagle area. One pre-collar drill hole was deepened 246 feet with core drilling and one core drill hole was completed from surface to a depth of 3,806 feet in the American Eagle area. Redhawk also drilled three core drill holes from surface totaling 10,975 feet.

Redhawk drilled twelve core holes totaling 3,800.4 feet from surface on the Copper Prince Breccias and three core holes from surface on the Globe Breccias totaling 1,220 feet in early 2008.

IMC updated the property mineral resources in October 2008. This estimate included resources in the Globe and Copper Prince breccias for the first time. The resources are documented in the IMC’s NI 43-101 Technical Report titled “Copper Creek 2008 Mineral Resource, Pinal Country, Arizona, USA, Technical Report” dated October 28, 2008 and filed on SEDAR on October 29, 2008.

Redhawk commissioned K D Engineering of Tucson to provide a scoping level economic study for the project in late 2009, based on the 2008 resource estimate. The NI 43-101 Technical Document for this resource is titled “Copper Creek Project 2,500 - 10,000 TPD Scoping Study”, dated March 12, 2010 and filed on SEDAR on May 12, 2010.

Redhawk conducted a district exploration program in 2010 and early 2011. This program targeted previously six undrilled areas outside the existing breccia and porphyry resources which Redhawk judged prospective for potentially higher grade mineralization in mafic volcanic and diabase host rocks. Encouraging copper intercepts in hole REX-10-047 west of the Keel and American Eagle areas suggested the possibility of extending these porphyry resources westward. Consequently, two additional angled holes were drilled from the same site. Drilling for the expanded eight-hole program totaled 32,871 feet. A key result of the district exploration program was that it showed for the first time the scale of the Copper Creek sulfide system. These holes expanded the footprint of known porphyry-style alteration and sulfide mineralization from approximately 5,500 to 13,000 feet in the northwest-southeast direction.

Starting in February 2011, Redhawk embarked on a 100,000-foot (30,000 meter) program of in-fill and step-out drilling intended to upgrade a significant portion of the American Eagle and Keel porphyry resources from the 'Inferred' category to the 'Measured and Indicated' category. In addition, the program was designed increase the size and confidence of the potential resource area connecting the Keel and American


SGS FORM No. A263a-2/14/13 !

! 16

Eagle porphyry resources. Both objectives were successfully achieved. The program included both vertical and angled core holes, and completed core drilling from a number of rotary pre-collars drilled in 2007 and 2008. As of the end of July 2012, 77 holes (plus a wedge from one hole) totaling 190,388 feet had been completed by Redhawk on the project.

At Redhawk's request, IMC updated property-wide breccia and porphyry resources in April-May 2012, based on drilling through end-March, 2012. This NI 43-101 compliant mineral resource estimate was announced in Redhawk's news release dated May 10, 2012, and forms the basis of the resource report titled “Copper Creek 2012 Mineral Resource Update, Pinal County, Arizona, USA – Technical Report”, dated June 25, 2012. This resource estimate considered potential open-pit extraction of a much larger scale resource.

Redhawk and its consultants evaluated the opportunities of developing Copper Creek as either a large open pit at mill rates of 100,000 to 120,000 tons per day or a large scale underground mine with mill rates in the 20,000 to 25,000 tons per day. To support the underground alternative, the Copper Creek mineral resource estimate was updated with any additional information to target an underground resource. The mineral resource estimate was revised for the Keel and American Eagle deposits based on 174 drill holes (343,799 feet) and the mineral resource for the breccia deposits is based on the mineral resource announced in the October 28, 2008 Technical Report. This resource estimate was announced in the Redhawk press release dated December 19, 2012 and is documented in the technical report titled “Copper Creek December 2012 Mineral Resource - Pinal Country, Arizona, USA - Technical Report”, dated January 31, 2013. This mineral resource is the basis of the current PEA.


SGS FORM No. A263a-2/14/13 !

! 17

7.0 GEOLOGICAL SETTING AND MINERALIZATION

The Copper Creek Property lies in the porphyry copper metallogenic province of the southwestern United States of America.

7.1 Regional Geology

The regional basement consists of a variety of Proterozoic rocks that lie on the southern flank of the Achean craton of North America. During the Mesozoic Era, back arc and foreland basin volcanic and sedimentary formations covered much of southern Arizona. This Mesozoic event was succeeded by a Laramide age (Late Cretaceous - Early Tertiary; 80-50 Ma.) volcanic-plutonic arc event accompanied by a major ENE-WSW compressional tectonic regime. This tectonism resulted in northwest-trending, basement sourced uplifted blocks and a widespread ENE-trending brittle structural fabric. The majority of porphyry copper deposits in the southwestern United States (including Copper Creek) date to this period of Laramide tectonism. Thick Galiuro Volcanics of mid-Tertiary age then covered the area and subsequent erosion has locally re-exposed the underlying copper-bearing bedrock. Figure 7.1 is a regional geologic map showing the geologic units by age.

The Copper Creek Property lies at the intersection of an ENE-trending belt of porphyry copper deposits, which include Ajo, Lakeshore, Silver Bell, San Manuel/Kalamazoo, Safford and Morenci, and a NNW-trending belt that includes Superior (Resolution), Ray, Christmas and Miami-Inspiration. Many of these deposits are tilted, faulted, and deeply eroded, but the Copper Creek porphyry system appears to be nearly upright and largely intact. Previous exploration and development have taken place on high level porphyry copper breccia pipes. Below these breccias, drilling has indicated an extensive porphyry copper-molybdenum mineralized system, which has been the focus of Redhawk’s exploration in recent years.

7.2 Copper Creek Geology

The Copper Creek district is centered on the Copper Creek granodiorite, the central of three Laramide granodiorite intrusions forming a northwest-oriented cluster on the west side of the Galiuro Mountains. The Copper Creek granodiorite was emplaced approximately 62 million years ago into Precambrian and Paleozoic sedimentary rocks, Late Precambrian diabase, and Cretaceous Glory Hole volcanics. The Copper Creek stock and adjacent Glory Hole volcanics have been intruded by a sequence of Laramide granodiorite, monzogranite, and quartz diorite porphyry plugs and dykes. The district is marked by over 400 hydrothermal breccia bodies, ranging from a few feet to several hundred feet across, which (like the porphyry bodies) are concentrated in two northwest-trending belts. Post-mineral Galiuro Volcanics cover all these rocks on the east and northeast. To the southwest, the district is bounded by a northwest-trending range-front fault which downdrops Tertiary Gila Conglomerate against the Laramide and older rocks. Figure 7.2 is a geologic map showing the mapped lithologies in the Copper Creek area.


SGS FORM No. A263a-2/14/13 !

! 18

Figure 7.1 - Regional Geology Map


SGS FORM No. A263a-2/14/13 !



SGS FORM No. A263a-2/14/13 !


Figure 7.2 - Local Geology - Mapped Lithologies


SGS FORM No. A263a-2/14/13 !


7.3 Mineralization

The Copper Creek mineralization consists of chalcopyrite, bornite, and chalcocite. Copper oxide minerals locally replace sulfides in the near surface supergene weathering environment. A chalcocite enriched, supergene zone forms the principal part of the Old Reliable deposit above the water table. Molybdenite occurs in sufficient amounts in the Childs-Aldwinkle breccia, the Old Reliable breccia, the Keel, and the American Eagle to be of economic interest. Gold and silver occur but were not systematically assayed by previous owners. Redhawk has re-assayed individual intervals to better understand the precious metal grade distribution where previous owners had assayed long composite runs and routinely assays for silver in drilling.

Near-surface mineralization at Copper Creek occurs predominately in the breccia bodies and to a lesser extent in high angle “early halo” veins. Sulfides are concentrated within the breccia matrix, and are dominated by chalcopyrite, bornite, and pyrite, with minor chalcocite in the upper parts of some pipes. High-grade mineralization in the breccias typically occurs as prominent masses and clots of copper minerals filling open spaces, or as sheeted veins along the edges of the pipes. Relative levels of copper, molybdenum, gold and silver vary considerably from breccia to breccia. Molybdenite Re-Os ages from the breccias yield mineralization ages between 62 and 57 million years.

Recent drilling indicates that the deeper American Eagle and Keel porphyry-style resources are connected and form a single copper-mineralized body at least 5000 ft long northwest-southeast, which is open in most directions. This mineralized body is controlled by a broad dome-shaped zone of common, subhorizontal to steep “early halo” quartz-sulfide veins. These veins display thin centerlines of quartz and sulfides, surrounded by much thicker halos of biotite, sericite, and abundant copper sulfides. The early halo vein zone is hosted mostly in Copper Creek granodiorite but appears centered on a cluster of syn-mineral granodiorite porphyry bodies. In parts of the American Eagle/Keel, many early halo veins are re-opened as semi-massive chalcopyrite-bearing veins that enhance the copper grade. Superimposed on the early halo vein zones are breccias and zones of intense quartz-sericite alteration, both of which tend to carry high-grade copper which further upgrades copper mineralization within the same volume of rock. A near-vertical set of early halo veins crosscuts and extends above the well mineralized dome-shaped early halo vein zone to the present surface, where outcropping veins are widespread and typically trend east-northeast. Sulfides in the American Eagle/Keel zone are zoned with depth, with pyrite-dominant mineralization near the surface transitioning into chalcopyrite-dominant rock in the better mineralized zones and increasing bornite content at depth.

Small lead and zinc prospects, primarily vein occurrences, occur toward the edges of the breccia-porphyry system.


SGS FORM No. A263a-2/14/13 !


8.0 DEPOSIT TYPES

The Copper Creek area contains multiple styles of Laramide copper-molybdenum-silver+/-gold deposits characterized by relatively high primary copper grades. Early development of the district dated from 1863, and focused on the exposed copper-rich (>>1% Cu) breccia bodies and peripheral silver-lead-zinc veins. In the 1960s and 1970s deeper drilling discovered porphyry-style sheeted and stockwork vein mineralization (~0.8% Cu) at depths between 1,200 to +4,000 feet in the American Eagle and Keel areas, beneath a small portion of the near-surface breccia cluster.

The breccias are clast- to matrix-supported and consist of angular to subrounded, pebble- to boulder-sized, commonly quartz-sericite altered fragments formed from the host wall rock. Matrices of most breccias are partially to completely filled with varying combinations of quartz, sulfide minerals, tourmaline, specularite, and minor rock flour. Breccias are known to persist over 3000 feet vertically. They terminate abruptly upward into lower grade material; the Mammoth pipe, the largest breccia in the current resource, is “blind” a mere 105 feet below the surface. Where drilling density is sufficient, the bottoms of the breccia pipes neck downwards into bodies of granodiorite porphyry. Over 90 percent of the mapped breccia bodies have not been drilled.

Evidence suggests that the breccia pipes were formed during implosive wall-rock collapse initiated by widespread porphyry magma withdrawal. As magma withdrew from the apex of a given intrusive body, gravity induced collapse initiated brecciation of the wall rock. The brecciation propagated to form the breccia bodies that are elongate in the vertical direction.

The American Eagle and Keel porphyry zones were previously classified as “hybrid porphyry” or “sheeted vein type” deposits due to their unusual vein styles. Recent work by Redhawk geologists has led to the recognition that the Copper Creek porphyry deposits are typical of the “early-halo type” of porphyry system described by Proffett (2009). In these deposits the dominant style of copper-bearing veins are early dark micaceous (EDM) veins or early potassic halos lining incipient fractures, rather than the A-type sugary quartz stockwork veins common in many porphyry systems. Well known early-halo type porphyry deposits include Butte, Montana, and Chuquicamata and Los Pelambres, Chile.


SGS FORM No. A263a-2/14/13 !


9.0 EXPLORATION

Exploration activities on the Copper Creek Property have spanned over 140 years by many different companies as noted in Section 6, History. The exploration work conducted since 1950 focused on the two different mineralization types: the breccia pipes or the deep porphyry deposits. The focus depended on the mining/exploration company doing the work with the larger, copper producing companies focusing on the large porphyry targets and the smaller companies (looking to develop production quickly) focusing on the breccia targets.

AMT mapped the surface expressions of breccia occurrences on the property and Redhawk is expanding on this work. Portions of an internal Redhawk memo are included below to provide an initial measure of the exploration potential on the property.

“Between the coordinates of 630,000 to 646,000 north and 394,000 to 410,000 east, a total of 358 breccia occurrences were counted, not including the Old Reliable, Childs-Aldwinkle or Mammoth. As best can be determined, 34 of the individual breccias have been drill tested on the Redhawk property, many with multiple holes. This figure includes only holes which appear to penetrate a specific breccia, and does not take into account roughly 30 holes scattered throughout the property which were not collared in (or directed towards) a breccia, but may well have intercepted the irregular subsurface continuation of one of the bodies. From this it follows that there are at least 324 of the mapped breccia occurrences on Redhawk property that have not been tested by drilling. With the acquisition of the PD claims and additional surface exploration, the number of mapped breccias now exceeds 400.”

If the heavily drilled area is excluded, then within the Redhawk property south of the PD claim block, there are 235 breccia bodies that have been tested by only ten drill holes. Many of these breccias have had surface samples collected and analyzed with about 590 samples taken in the breccias. The compilation of this data by Redhawk reveals:

“The western breccia belt was found to have 15 samples anomalous in copper (>200 PPM) from 14 breccias, of which one has been drill tested. One anomalous untested molybdenum (>40 PPM) sample was noted from breccia sampling in the western belt. In the southern area are 37 breccia samples anomalous in copper from roughly 26 breccia bodies, with one of these having a drill test. Ten breccia samples in this area are anomalous in molybdenum from 8 bodies, again with one having a drill test. Eight of the breccia bodies from the southern area have coincident anomalous copper and molybdenum values, and one of these has been drill tested. In the northeast are there are only about 14 breccia samples, with two of these anomalous in copper and one anomalous in molybdenum, with no drilling to test the anomalous samples.”

Work by Redhawk since 2009 has resulted in determining that the Copper Creek mineralization system is an “early halo” style characterized by a large percentage of


SGS FORM No. A263a-2/14/13 !


“early dark micaceous” (EDM) and later “D” type veining hosting the vast majority of copper and molybdenum values in the “porphyry” mineralization. The abundant breccia bodies appear to be a late phase of mineralization. The EDM and D style veins are identifiable in road cuts, drainages, and sometimes in outcrops. The EDM and D style veins are being mapped on the surface and show a strong correlation with the “footprint” of the porphyry mineralization area. Zones of more intense EDM and D veining on the surface are most commonly ENE trending and have very steep to vertical dips and are target areas for developing open pit mineable grades above the more massive zones of porphyry mineralization at deeper levels. The deeper porphyry mineralization is dominated by the EDM and D veining but dips of the veining is strongly sub-horizontal to moderate angles rather than the steep to vertical dips at shallower levels.

Detailed drill hole logging has resulted in determining a series of porphyritic intrusions that intrude the main granodiorite body that hosts the vast majority of the porphyry style mineralization. Crosscutting relationships between the porphyritic intrusions and various veining episodes are being used to develop a full three dimension geologic model in the resource areas. The Redhawk geologists have now identified five unique porphyries. In these porphyries there are early halo and the later D vein systems, and there are instances where the later D veins have been cut by early halo veins of a later porphyry intrusive pulse.

During 2010 and 2011 six widely spaced drill holes were drilled on exploration targets within the district. The holes were designed to test various target concepts and most of the test holes were 1 to 2 kilometers from the edge of the 2008 deep porphyry resource of the Keel and American Eagle. Five of the six holes encountered significant grades of copper and molybdenum and alteration and veining styles consistent with that found in the resource area. The drilling and the presence of the breccia bodies over a large area indicate a large hydrothermal system that is largely under explored.

Drilling during 2011 and 2012 focused on drill targeting to move as much of a large “inferred” resource tonnage in the American Eagle/Keel into the higher quality “indicated” and/or “measured” resource categories. The drill program also made a strong connection between the American Eagle and Keel mineralization showing the area to be one continuous mass of similar mineralization.

Recent work by the Redhawk geologists has included the mapping of the intensity and location of many trace elements and their relationship to the copper, molybdenum and precious metal mineralization. The Keel-American Eagle deposit has the most drill hole information with many of the drill intervals assayed for a full 32 element suite in addition to copper, molybdenum, silver and gold assays. The assay results from surface sampling over the Keel-American Eagle deposit were added to database. Levels and sections of several of the elements show strong zonation patterns in relation to the porphyry. The results of this work show that the more volatile elements tend to move differently away from the porphyry where the base metals are


SGS FORM No. A263a-2/14/13 !


concentrated. There appears to be a sequence of these elements with distance from the porphyry.

Using this model, the patterns of the trace elements were studied in more sparsely drilled and sampled areas, including areas of only shallow drilling. The mapping of the intensity and sequence of the trace elements has indicated the potential of several other large porphyry systems to the west and north of the current Keel-American Eagle porphyry.

Another interesting result of this trace element mapping is the occurrence of the trace elements which usually appear above the porphyry system being found in similar intensity below the Keel-American Eagle porphyry in holes which extended to depth below the porphyry. Using the trace element model, this leads one to believe that the potential exists for another porphyry below the Keel-American Eagle in a stacked system.

Additional drilling is needed to test this hypothesis of deeper porphyries below the Keel-American Eagle and in other locations within the Copper Creek claim block. If successful, the ability to target the location of other porphyries using this metal zonation model can lower the cost and speed up the exploration for and delineation of other porphyries which may exist on the un-tested areas of the Copper Creek claim block, thus increasing the mineral resource of Copper Creek.


SGS FORM No. A263a-2/14/13 !


10.0 DRILLING

Exploration drilling has been conducted by most of the previous owners of the Copper Creek property, and Redhawk commenced drilling during the fourth quarter 2006. A summary of the number of holes and footage by company is summarized in Table 10.1 which totals 551 holes for 659,521 ft. Table 10.2 summarizes the drilling by drill type and this total will exceed the 551 number of holes because some holes were drilled using more than one drilling method. Since the June 2012 mineral resource estimate and up to the cutoff date for the current mineral resource update, Redhawk has added 6 holes for an approximately 26,861 feet of drilling information to the Copper Creek database.

There are six mineralized areas which are of interest at Copper Cheek and are the combinations of the breccia zones (Globe, Copper Prince, Old Reliable, Childs-Aldwinkle, and the Mammoth) with the deeper porphyry mineralization (Keel, which is below the Mammoth breccia and the American Eagle, to the southeast of the Keel). Figure 10.1 shows the drill holes and the limits of the mineralized areas. Table 10.3 summarizes the number of holes, footage and number intervals plus intervals with assays for copper, moly, and silver within each of the mineralized area limits shown on the map. The total number of holes on this table exceeds the 551 total because some holes cross from one area to another.

Table 10.1 Drill Hole Statistics by Company

Data Base Code Company Date

Number of Holes

Total Drilling, ft

Assayed for Copper

Assayed for Molybdenum

13 Calumet & Arizona 1914 14 5,409 3,232 0

2 Bureau of Mines 1942-3 31 2,931 2,753 0

7 Siskon 1956-58 25 4,027 3,914 760

3 Bear Creek Mining 1959-62 15 29,085 17,163 9,944

6 Newmont 1966 22 30,258 27,274 17,347

10 Occidental 1968-70 49 9,219 8,516 3,340

9 Ranchers 1971 3 785 785 0

5 Magma 1971-72 38 94,271 91,115 30,428

11 Exxon 1971-72 21 73,529 72,809 72,281

12 Inspiration (?) 1973 6 746 746 81

15 Phelps Dodge 1972-74 9 25,447 24,408 30

1 AMT 1995-2001 238 192,407 103,521 90,921

16 Redhawk 2006 on 78 190,388 111,109 111,109


SGS FORM No. A263a-2/14/13 !


Others 2 1,020 0 0


SGS FORM No. A263a-2/14/13 !


Table 10.2 Drilling Statistics by Drilling Type

Diamond Core

Reverse Circulatio

nRotary/

PercussionDrifts/ Raises

Not Classifie

d Total

Number of holes 370 75 95 56 11 607

Total drilling, ft 524,818 35,787 89,131 3,056 6,730 659,521

Assayed for copper, ft 420,437 29,748 57,084 2,878 5,660 515,807

Percent complete 80 83 64 94 84 78

Total assay intervals 46,079 3,009 5,868 577 566 56,099

Average assay interval, ft 9.1 9.9 9.7 5.0 10.0 9.2

Assayed for molybdenum, ft 318,690 17,758 43,477 110 4,700 384,734




Table 10.3 Drilling Statistics by Mineralized Area

Mineralized AreaNumber of Holes

Total Drilling,

feet

Number of

Intervals

Intervals Assayed

for Copper

Intervals Assayed for

Molybdenum

Intervals Assayed for Silver

Globe 35 28,198 2,145 2,085 1,101 582

Copper Prince 71 40,062 3,770 3,704 1,248 394

Old Reliable 141 42,417 4,780 4,623 2,198 302

Childs-Aldwinkle 86 78,950 4,103 3,886 3,125 1,032

Mammoth-Keel 85 141,860 11,636 11,392 11,151 5,854

American Eagle 94 242,688 22,851 22,552 17,483 11,294

Outside 85 85,347 7,942 7,854 5,261 4,254

Total 659,521 57,227 56,096 41,567 23,712


SGS FORM No. A263a-2/14/13 !


Figure 10.1 Drill Hole Locations

Mineralized Areas 1 = Globe 2 = Copper Prince 3 = Old Reliable 4 = Chiles-Aldwinkle 5 = Mammoth – Keel 6 = American Eagle


SGS FORM No. A263a-2/14/13 !

N

5

1

6

3 4

2


11.0 SAMPLE PREPARATION, ANALYSES AND SECURITY

11.1 Prior to Redhawk

Redhawk does not have any record of the methods used for core logging and sampling from drilling that predates AMT’s involvement with the property (AMT optioned the property in 1995). The earlier drilling was done by several major mining and resource industry companies and it is assumed that they used industry standard practices at the time. Review of the geologic logs, down hole survey records and assay certificates that Redhawk has on file confirms that data was collected and handled with good practices.

As the various companies participated in the exploration of Copper Creek, assay checks were often done on previous assay work as part of the new company’s due diligence. This check assay work was reviewed by IMC and is presented in Section 14 of the IMC report “Copper Creek 2008 Mineral Resource, Pinal County, Arizona, USA, Technical Report” dated October 28, 2008 and posted on SEDAR.

AMT started work at Copper Creek in 1995. There are no written records of its sampling methods and data handling, but verbal discussions by Joe Sandberg of Redhawk with Claus Wiese, a former employee of AMT, confirmed that the procedures followed industry standards. In summary, the drill core or cuttings were logged for geologic and geotechnical information, core was split by either sawing or mechanical splitting, and half core was sent for assay. The particulars of the sample split for the RC drilling is not known.

Redhawk has spent much time and effort organizing and reviewing the data from the previous work since it took over the property in 2005. A summary of its activities and chain of custody of the information is included in Section 13 of the IMC report mentioned above. Redhawk has consolidated the drill core, cuttings and assay pulps into one storage facility under its control. Redhawk has re-logged a majority of the core to both confirm the previous logging and to evaluate new theories. Redhawk found the core from several of the holes drilled by AMT late in its property ownership had not been split or assayed. Redhawk has sampled and assayed these holes and updated the drill hole data base.

11.2 Redhawk Sample Handling Procedures

All drill core (or RC samples) is either picked up on the drill site by Redhawk personnel and brought directly to Redhawk’s secure facility in San Manuel, AZ, or is delivered directly to Redhawk’s facilities by the drilling crew. Redhawk’s geologists log the core and mark the core for sampling. Redhawk personnel saw or split the core as per the geologist’s marked intervals and split lines. One half of the split core is placed in marked bags and delivered by Redhawk personnel directly to Jacobs Assay in Tucson, AZ for sample prep. The second half of the core is stored in the Redhawk core storage


SGS FORM No. A263a-2/14/13 !


facilities for reference. IMC has observed the logging and splitting practices during its 2008 visits and found them to meet industry standards. Redhawk has confirmed that the practices have not changed.

11.3 Current Sample Preparation Procedures

The split core is delivered to Jacobs Assay for preparation of a pulp sample for assaying. The procedure is documented below from a verbal conversation with Michael Jacobs.

▪ On delivery, the samples are checked against the list provide by Redhawk, and then logged into the lab records in the same order as on the Redhawk list.

▪ The entire sample is processed through a jaw crusher, three times to reduce the size to a nominal minus 1/2 inch.

▪ The jaw crusher is brushed off and blown clean with compressed air between each sample.

▪ The sample is mixed, pan to pan, six times; then poured through a Jones splitter. One quarter of the sample is retained for further processing and the remaining three quarters is placed back in the bag as the reject sample. These are retained for future use. The Jones splitter is cleaned with compressed air between each sample.

▪ The one quarter split of the sample goes through a second jaw crusher and roller (one pass), mixed, and further crushed if needed to reach minus 10 mesh size.

▪ The sample passes through a Jones splitter to get a 250 gram split, which is pulverized to minus 150 mesh. The Jones splitter is cleaned and the pulverizer is cleaned with silica sand after each sample.

▪ The pulverized 250 gram sample is mixed on rolling cloth 25 times from each corner, and then placed into the pulp sample bag for shipment to the assay laboratory.

▪ The rejected minus 10 mesh material that was not pulverized is placed in a separate bag that is placed in the bag of the reject material from the initial split.

The prepared pulps are shipped directly by public carrier from Jacobs to ALS Chemex Laboratory in Reno, Nevada for assay. Reject samples designated for


SGS FORM No. A263a-2/14/13 !


metallurgical work by Redhawk personnel are sorted by Jacobs and delivered to Metcon in Tucson, AZ.

Jacobs prepares a second pulp from the sample reject material for every tenth sample and those pulps are given to Redhawk personnel or delivered back to the Redhawk office. Once several jobs of reject check pulps are collected, these are boxed and shipped to ALS in Reno for assay.

In 2007, Jacobs prepared five standards (one to three 5 gallon buckets each) of various copper assay levels from Copper Creek mineralized material. Either 5 or 10 samples from each standard created were sent to three separate laboratories for copper, molybdenum and silver assaying. The results from each laboratory were compared and an average value for the standard created for copper, molybdenum and silver. One of the five standards is selected at random and inserted in with the pulps sent to ALS for assaying. The standards are inserted at the rate of about 1 in 10. The standards vary from 0.014 to 5.678 percent copper and from 0.00032 to 0.0092 percent molybdenum. In late 2012 a sixth standard was prepared and 10 samples were sent to 5 certified labs for assay of copper, molybdenum, and silver. The results from each lab were compared and an average value for the standard created at 0.745 percent copper, 0.0285 percent molybdenum, and 3.13 ppm silver. A blank standard has been prepared and is being readied for use in future work.

11.4 Assay Procedure

Redhawk samples were analyzed by ALS Chemex in Reno, Nevada. The samples are assayed for copper and molybdenum using standard four acid digestion assay techniques. Significantly mineralized intervals were then re-assayed for gold using 30 gram charge fire assay and AAS and multi-element ICP (48 elements) to get silver values, copper and molybdenum check values, and other elements for geochemical and metallurgical information. Copper values from the ICP exceeding 10,000 ppm were re-assayed using an additional "ore grade" assay method. All drilling starting in 2010 is assayed by ALS for multi-element methods with copper values exceeding 5,000 ppm or, later in the drilling program, 2,000 ppm being re-assayed by AA 62 and values exceeding 10,000 ppm are re-assayed using an additional "ore grade" assay method.

Approximately one in 12 to 14 original pulps is selected for check assay by Redhawk. The list of pulps for check assay is sent to ALS, which retrieves the pulp samples and these samples are transferred to the Inspectorate Lab in Reno. The check samples are assayed for copper, molybdenum and silver with the results provided to Redhawk.


SGS FORM No. A263a-2/14/13 !


11.5 Independent QAQC Review

Redhawk commissioned an independent review of the project quality assurance and quality control (QAQC) program in October 2012. This review was conducted by industry recognized expert Dr. Jeffrey Jaacks (Geochemical Applications International Inc.). Dr. Jaacks examined all controls and procedures related to sample preparation and security, internal standard samples, duplicates, inter-lab check analyses, and ongoing QAQC monitoring to ensure adherence to industry best practices. Dr. Jaacks' reports to Redhawk dated 15, 16, and 30 October 2012 indicate the company passed the review without any issue. Dr. Jaacks made minor recommendations to raise QAQC protocols used on the project to an even higher standard.


SGS FORM No. A263a-2/14/13 !


12.0 DATA VERIFICATION

This section is taken from the IMC technical report titled “Copper Creek 2012 Mineral Resource Update, Pinal County, Arizona, USA - Technical Report” dated June 25, 2012 (posted on SEDAR). Redhawk has continued to run assay checks, standards and blanks and this data is being supplied to IMC who is in the process of updating its findings.

12.1 Pre-2010 Drilling

Check assay and other relevant results for the earlier drilling conducted by Redhawk and preceding companies at Copper Creek are discussed in Section 14 of the NI43-101 Technical Report entitled “Copper Creek 2008 Mineral Resource” dated October 28, 2008 and posted on SEDAR. The review of the available assay data base used for the 2008 Mineral Resource was divided into two parts: review of the assay data itself in the form of check assays run by previous owners and spot check the Redhawk data base as it was provided to IMC against original data in the form of assay certificates, geologic logs and survey notes. IMC found the data to be acceptable for the development of a resource estimate and notes some areas that require additional attention as the project continues.

12.2 2010 - 2012 Redhawk Drilling - Summary

This review summarizes the results of the QAQC and check assaying program on the Redhawk drilling conducted from 2010 through 2012. The data base assays for the 2010-12 Redhawk drilling were run by ALS on pulps prepared by Jacobs Assay. The QAQC data supplied to IMC consist of assays run by ALS on standards, check assays run by Inspectorate on Jacobs pulps, assays run by ALS on a second Jacobs pulp from the sample coarse reject and assays run by METCON on pulps prepared by METCON. Based on its analysis of these data IMC's conclusions are:

▪ ALS's assaying procedures for copper, molybdenum and silver are acceptable.

▪ The only assays that act as a complete check on Jacob's sample preparation procedures are those run by METCON on METCON pulps, and the fact that these assays are biased up to 17 percent low relative to the ALS data base assays for copper indicates a possible sample preparation problem at METCON or at Jacobs. Jacobs did prepare a second pulp from the coarse reject material for selected samples and this pulp was assayed by ALS and falls within acceptable ranges. IMC believes that the problem most likely occurred at METCON, but the possibility that it occurred at Jacobs cannot be discounted at this time. As a result, IMC believes that the 2010-12 Redhawk assays are conditionally acceptable until Jacobs' sample preparation procedures are validated. This validation needs to be done prior to any future


SGS FORM No. A263a-2/14/13 !


resource updates. Redhawk has sent a batch of 73 coarse reject samples (from intervals where the pulp was prepared by Jacobs) to ALS for sample preparation and assay. Redhawk has commented to IMC that this check work came back fine, and IMC needs to follow up on the results.

▪ To do this it is recommended to ship every 20th reject sample from the post-2008 drilling (approximately 600 samples) to a check laboratory, have this laboratory prepare pulps from the rejects and assay them with the same procedures as used by ALS. If the assays compare acceptably with the data base assays, the Jacobs sample preparation procedures will be validated. (IMC recommended in Section 14.4 of the October 2008 Technical Report that future check assays be run on rejects rather than pulps to obtain a check on both analytical and sample preparation procedures.)

▪ Consideration should be given to adding a higher-grade molybdenum standard and a silver standard. Since the June 2012 technical report, Redhawk has added a 0.028 percent molybdenum standard to its suite of standards. A silver standard is also being run.

▪ Blank samples should be inserted in the sample stream as part of the QAQC program. Redhawk has created a blank sample for use once drilling resumes. At that time, the round robin assay work from a group of assay labs will be done to verify the assay of the blank standard.

12.3 Assays on Standards

The results of the ALS Chemex assays on five copper - molybdenum standards are summarized in Tables 12.1 and 12.2. A standard was inserted approximately once every 13 samples (1,625 standards in 12,725 total assays). The five standards were prepared from Copper Creek drill samples by Jacobs and expected grades were determined by assaying the standards at three assay laboratories and taking the mean of all the assays. There are no silver standards.

Table 12.1 Assays on Copper Standards

Standard Expected Grade, %

Number of Readings,

%

Mean, % Standard Deviation,

%

Median, %

% Mean/ Expected

% Median/

Expected

1 0.042 319 0.038 0.010 0.036 -10 -14

2 0.969 470 0.948 0.048 0.951 -2 -2

3 0.555 410 0.538 0.043 0.534 -3 -4

4 5.678 347 5.517 0.374 5.559 -3 -2

5 0.014 79 0.015 0.006 0.014 +7 0


SGS FORM No. A263a-2/14/13 !


The mean copper grade of the standard assays is 2 to 4 percent lower than the expected grade for standards 2, 3 and 4 and the mean molybdenum grade is 5 to 7 percent lower than the expected grade for standards 2 and 3. The grades of the other standards are too low to allow a meaningful mean/expected grade comparison to be made. (Median values are shown in the tables because the scatter of the standard assays makes it difficult to determine whether the outlier points are a result of assay variability, poor homogenization of the standard, sample mislabeling or data entry errors. Under these circumstances the median might be considered a more meaningful estimate of the mean of the standard assays.)

The results of the standard assays are summarized in Figures 12.1 through 12.10. There is no indication of any significant variations in grade with time for any of the standards.

IMC considers these results acceptable. However, QAQC programs commonly include assays run on blanks and it would therefore be desirable to insert blanks into the sample stream. (Standard 5 could be replaced by blanks). Adding a higher-grade molybdenum standard in the +/- 0.05 percent range and a silver standard in the +/- 1 g/t range should also be considered.

Table 12.2 Assays on Molybdenum Standards

Standard Expected Grade, %

Number of

Readings

Mean, % Standard Deviation,

%

Median, %

% Mean/ Expected

% Median/

Expected

1 0.0030 319 0.0031 0.0024 0.0028 +3 -20

2 0.0075 470 0.0073 0.0018 0.0070 -3 -7

3 0.0092 410 0.0088 0.0021 0.0087 -4 -5

4 0.0010 347 0.0014 0.0004 0.0014 +40 +40

5 0.0003 79 0.0004 0.0001 0.0004 +33 +33


SGS FORM No. A263a-2/14/13 !


Figure 12.1 - Standard 1, Copper, Expected Value 0.042%

X-Axis = Days from October 1, 2010 (Plot Ends April 8, 2012)


X-Axis = Days from October 1, 2010 (Plot Ends June 8, 2012) Figure 12.3 - Standard 3, Copper, Expected Value 0.555%


SGS FORM No. A263a-2/14/13 !

0

0 . 0 2

0 . 0 4

0 . 0 6

0 . 0 8

0 . 1

0 . 1 2

0 . 1 4

0 1 0 0 2 0 0 3 0 0 4 0 0 5 0 0 6 0 0

0

0 . 1

0 . 2

0 . 3

0 . 4

0 . 5

0 . 6

0 . 7

0 . 8

0 . 9

1

1 . 1

1 . 2

0 1 0 0 2 0 0 3 0 0 4 0 0 5 0 0 6 0 0 7 0 0


X-Axis = Days from October 1, 2010 (Plot Ends June 8, 2012)


X-Axis = Days from October 1, 2010 (Plot Ends May 28, 2012)

Figure 12.5 - Standard 5, Copper, Expected Value 0.014% X-Axis = Days from April 6, 2012 (Plot Ends June 8, 2012)


SGS FORM No. A263a-2/14/13 !

0

0 . 1

0 . 2

0 . 3

0 . 4

0 . 5

0 . 6

0 . 7

0 . 8

0 . 9

1

1 . 1

1 . 2

0 1 0 0 2 0 0 3 0 0 4 0 0 5 0 0 6 0 0 7 0 0

0

1

2

3

4

5

6

7

0 1 0 0 2 0 0 3 0 0 4 0 0 5 0 0 6 0 0


Figure 12.6 - Standard 1, Molybdenum, Expected Value 0.0030%

X-Axis = Days from October 1, 2010 (Plot Ends April 8, 2012)



SGS FORM No. A263a-2/14/13 !

0

0 . 0 1

0 . 0 2

0 . 0 3

0 . 0 4

0 . 0 5

0 . 0 6

0 . 0 7

0 . 0 8

0 . 0 9

0 . 1

0 1 0 2 0 3 0 4 0 5 0 6 0 7 0

0

0 . 0 1

0 . 0 2

0 . 0 3

0 . 0 4

0 . 0 5

0 1 0 0 2 0 0 3 0 0 4 0 0 5 0 0 6 0 0


X-Axis = Days from October 1, 2010 (Plot Ends June 8, 2012)


X-Axis = Days from October 1, 2010 (Plot Ends June 8, 2012) Figure 12.9 - Standard 4, Molybdenum, Expected Value 0.0010%

X-Axis = Days from October 1, 2010 (Plot Ends May 28, 2012)


SGS FORM No. A263a-2/14/13 !

0

0 . 0 0 5

0 . 0 1

0 . 0 1 5

0 . 0 2

0 . 0 2 5

0 . 0 3

0 . 0 3 5

0 . 0 4

0 1 0 0 2 0 0 3 0 0 4 0 0 5 0 0 6 0 0 7 0 0

0

0 . 0 0 5

0 . 0 1

0 . 0 1 5

0 . 0 2

0 . 0 2 5

0 . 0 3

0 . 0 3 5

0 . 0 4

0 1 0 0 2 0 0 3 0 0 4 0 0 5 0 0 6 0 0 7 0 0



X-Axis = Days from April 6, 2012 (Plot Ends June 8, 2012)


SGS FORM No. A263a-2/14/13 !

0

0 . 0 0 1

0 . 0 0 2

0 . 0 0 3

0 . 0 0 4

0 . 0 0 5

0 . 0 0 6

0 . 0 0 7

0 . 0 0 8

0 . 0 0 9

0 . 0 1

0 1 0 0 2 0 0 3 0 0 4 0 0 5 0 0 6 0 0 0

0 . 0 0 1

0 . 0 0 2

0 . 0 0 3

0 . 0 0 4

0 . 0 0 5

0 1 0 2 0 3 0 4 0 5 0 6 0 7 0


12.4 Check Assay Comparisons

The results of the check assay comparisons for copper are summarized in Table 12.3. Copper represents the bulk of the economic value of the Copper Creek resource, so the copper comparisons are the most important.

According to these results a Jacobs pulp was sent to Inspectorate for check assay approximately once every 23 samples (562 assays in 12,725 total assays) and a second pulp prepared from the coarse reject by Jacobs was re-assayed by ALS approximately once every 43 samples (295 assays in 12,725 total assays). The METCON assays were run on pulps prepared separately from Jacobs rejects by METCON.

IMC generally considers check assay results to be acceptable when the mean grade of the check assays is within 5 percent of the mean grade of the data base assays and/or when the gradient of a QQ-plot, which compares the grade distributions of the two data sets rather than the mean grades, falls within 5 percent of a 1:1 line. This criterion is met for the ALS versus Inspectorate copper comparison. However, because both the ALS data base and the Inspectorate check assays were run on Jacobs pulps the results act to verify only ALS's analytical procedures.

The criterion is also met for the ALS versus ALS duplicate assay copper comparison. The ALS duplicate assays were run on second pulp prepared by Jacobs and therefore acts as a check only on analytical procedures. If there were any problems with Jacobs' sample preparation procedures the duplicate comparisons would not identify them.

The METCON assays are the only assays that were run on pulps prepared by a lab other than Jacobs and are therefore the only ones that act as an independent check on Jacobs' sample preparation procedures. The 5 percent criterion is met in the METCON (ICP) case but not in the two METCON AA cases, which show METCON

Table 12.3 ALS Database Assays (on Jacobs Pulps) Versus Check Assays for Copper

Check Lab Pulps from Number ALS mean, %

Check mean, %

Check/ALS %

QQ-plot %

Inspectorate Jacobs 562 0.366 0.379 +3.6 +3.0

ALS duplicates Jacobs 295 0.323 0.306 -5.2 -3.6

METCON (AA) METCON 97 0.570 0.470 -17.5

METCON (AA2)

METCON 61 0.545 0.490 -10.1

METCON (ICP)

METCON 61 0.545 0.542 -0.6


SGS FORM No. A263a-2/14/13 !


biased 10 to 17 percent low relative to ALS. These results suggest a sample preparation or assaying problem at METCON, but the possibility that the Jacobs sample preparation procedures are biasing the ALS copper assays high is being investigated by having ALS create new pulps from the coarse rejects to check the Jacobs sample preparation.

The results of the check assay comparisons for molybdenum and silver are summarized in Tables 12.4 and 12.5. These comparisons are less important because molybdenum and silver are byproduct minerals that contribute only a small proportion to the economic value of the Copper Creek resource.

The Inspectorate and ALS duplicate comparisons for molybdenum and silver also generally meet criteria for acceptance but again the ALS versus METCON comparisons show apparent biases, although in the opposite sense to copper.

The results of the check assays comparisons are shown in the following Figures:

Figures 12.11 through 12.15: XY scatterplots, copper Figures 12.16 and 12.17: QQ-plots, ALS/Inspectorate and ALS/ALS duplicate cases Figures 12.18 through 12.22: XY scatterplots, molybdenum

Table 12.4 ALS Database Assays (on Jacobs Pulps) Versus Check Assays, Molybdenum

Check Lab Pulps from Number ALS mean, % Check mean, %

Check/ALS %

Inspectorate Jacobs 562 0.0078 0.0080 -2.5

ALS duplicates Jacobs 294 0.0076 0.0073 +4.1

METCON (AA) METCON 97 0.0128 0.0109 +17.4

METCON (AA2)

METCON 61 0.0055 0.0050 +10.0

METCON (ICP)

METCON 61 0.0055 0.0048 +14.5

Table 12.5 ALS Database Assays (on Jacobs Pulps) Versus Check Assays, Silver

Check Lab Pulps from Number ALS mean, ppm

Check mean, ppm

Check/ALS %

Inspectorate Jacobs 562 1.09 1.17 +7.3

ALS duplicates Jacobs 294 1.08 1.07 -0.9

METCON METCON 97 1.77 2.02 +14.1


SGS FORM No. A263a-2/14/13 !


Figures 12.23 through 12.25: XY scatterplots, silver


SGS FORM No. A263a-2/14/13 !


! Figure 12.11 - ALS Database Assays vs. Inspectorate Check Assays, Copper


SGS FORM No. A263a-2/14/13 !


!Figure 12.12 - ALS Database Assays vs. ALS Assays on Second Pulp, Copper


SGS FORM No. A263a-2/14/13 !


!Figure 12.13 - ALS Database Assays vs. METCON AA Check Assays, Copper


SGS FORM No. A263a-2/14/13 !


!Figure 12.14 - ALS Database Assays vs. METCON AA2 Check Assays, Copper


SGS FORM No. A263a-2/14/13 !


!Figure 12.15 - ALS Database Assays vs. METCON ICP Check Assays, Copper


SGS FORM No. A263a-2/14/13 !


!Figure 12-16 - ALS Database Assays vs. Inspectorate Check Assays,

Copper, QQ-Plot, 99.8% of Data


SGS FORM No. A263a-2/14/13 !


!Figure 12.17: ALS Database Assays vs. ALS Assays on Second Pulp,

Copper, QQ-Plot, 99.5% of Data


SGS FORM No. A263a-2/14/13 !


!Figure 12.18 - ALS Database Assays vs. Inspectorate Check Assays, Molybdenum


SGS FORM No. A263a-2/14/13 !


!Figure 12.19 - ALS Database Assays vs. ALS Duplicate Assays, Molybdenum


SGS FORM No. A263a-2/14/13 !


!Figure 12.20 - ALS Database Assays vs. METCON AA Check Assays, Molybdenum


SGS FORM No. A263a-2/14/13 !


!Figure 12.21 - ALS Database Assays vs. METCON AA2 Check Assays, Molybdenum


SGS FORM No. A263a-2/14/13 !


!Figure 12.22 - ALS Database Assays vs. METCON ICP Check Assays, Molybdenum


SGS FORM No. A263a-2/14/13 !


!Figure 12.23 - ALS Database Assays vs. Inspectorate Check Assays, Silver


SGS FORM No. A263a-2/14/13 !


! Figure 12.24 - ALS Database Assays vs. ALS Assays on Second Pulp, Silver


SGS FORM No. A263a-2/14/13 !


! Figure 12.25 - ALS Database Assays vs. METCON Check Assays, Silver

13.0 MINERAL PROCESSING AND METALLURGICAL TESTING


SGS FORM No. A263a-2/14/13 !


The following is a discussion of the metallurgical testing and mineral processing aspects of the proposed project.

13.1 Metallurgical Testwork Discussion

13.1.1 Bulk Copper Molybdenum Flotation Test Program

The previous property owners of the Copper Creek property had metallurgical tests conducted at Mountain States R&D International, Inc. (MSRDI). Much of the test work evaluated heavy media separation or heavy media separation followed by flotation as processing techniques. Results from conventional flotation tests, described in the “Report of Metallurgical Testing on Copper Sulfide Resources of the Copper Creek Property” by MSRDI dated 3 January 1997 are summarized below. Additional work on other mineralized zones, conducted in 2008 is described in the “Copper Creek Project Preliminary Flotation Study” by METCON Research (METCON) dated November 2008. Results from this work are also summarized below.

In the first test series conducted by MSRDI, rougher flotation tests were conducted on four different composite samples from the Copper Creek Property. Three of the composites were from the Childs-Aldwinkle area and the fourth was from the Creek Breccia. These selected composite samples represented three copper grades: low (about 0.5 percent copper); average (1.5 to 2.5 percent copper); and high (plus 3 percent copper) for each mineralization type. The tests were planned to evaluate the effect of particle size (grind) on the recovery of copper and molybdenum.

Test conditions for the series are noted below:

▪ Lime = 3.5 lb/t pH = 10.8 - 11.2 ▪ S-5741 = 0.01 lb/t ▪ Nalco 9743 = 0.017 lb/t ▪ Z-11 = 0.002 lb/t ▪ Fuel Oil = 0.023 lb/t

Results from the test series are summarized in the following Table 13.1.


SGS FORM No. A263a-2/14/13 !


Table 13.1 Effect of Grind on Copper and Molybdenum Recovery in MSRDI Tests

!

As shown in Table 13.1, the copper recovery and rougher concentrate grades realized were both excellent. Generally, the effect of finer grind was to increase copper recovery. The effect of grind on molybdenum recovery is not as clear. The target grind P80 equal to 140 microns was selected for the remainder of the test program.

The second series of tests conducted by MSRDI, summarized in Table 13.2, was run to evaluate cleaning the rougher concentrate produced. The rougher concentrate was not reground, and excellent cleaner concentrate grades, averaging over 40 percent copper, were achieved.

Table 13.2 MSRDI Concentrate Cleaning Summary

!

Sample Test GrindIdentification Cu, % Mo, % No. Cu, % Mo, % P80* Cu, % Mo, % Cu, % Mo, % Cu Mo

(CA) CP & PY BM-1 0.57 0.0140 165 25.25 0.510 0.015 0.0026 97.42 81.51low grade BM-4 0.58 0.0157 135 25.55 0.550 0.017 0.0038 97.11 76.40

(CA) CP & BN BM-2 1.59 0.0080 170 35.73 0.146 0.146 0.0022 91.17 73.68medium grade BM-5 1.51 0.0099 145 36.04 0.173 0.032 0.0029 97.97 71.92(CA) CP & BN BM-3 3.44 0.0127 175 41.43 0.103 0.19 0.0050 94.90 63.76

high grade BM-6 3.40 0.0132 135 39.73 0.122 0.041 0.0032 98.90 77.89(CA) BN & CP BM-7 0.50 0.0149 170 21.28 0.470 0.022 0.0043 95.75 71.80

low grade BM-13 0.57 0.0128 150 25.65 0.422 0.017 0.0038 97.06 70.84(CA) BN BM-8 2.29 0.0237 195 43.57 0.364 0.048 0.0052 98.02 79.21

high grade BM-14 2.19 0.0187 165 41.59 0.294 0.025 0.0035 98.92 82.22(CA) BN BM-15 1.86 0.0290 175 37.77 0.540 0.029 0.0030 98.51 90.15

medium grade BM-9 1.82 0.0309 155 36.04 0.520 0.022 0.0052 98.85 84.03(CA) MO & CP BM-10 0.68 0.0463 200 24.94 1.450 0.021 0.0083 96.98 82.53

low grade BM-16 0.71 0.0506 150 24.48 1.540 0.015 0.0072 97.94 86.16(CA) MO & BN BM-17 1.96 0.1754 195 42.76 3.720 0.021 0.0069 98.98 96.25medium grade BM-11 1.75 0.1900 140 37.77 3.940 0.018 0.0096 99.02 95.19(CA) MO & CC BM-18 2.36 0.0745 210 36.24 1.040 0.127 0.0110 94.94 86.15

high grade BM-12 2.49 0.0960 145 36.34 1.180 0.097 0.0880 96.37 81.40(VIX) CP & PY BM-19 0.86 0.0030 175 22.19 0.023 0.015 0.0022 98.32 29.29

low grade(VIX) CP & PY BM-20 1.65 0.0031 165 25.14 0.012 0.015 0.0025 99.15 24.99 medium grade(VIX) CP & PY BM-21 2.90 0.0035 160 26.26 0.011 0.015 0.0026 99.54 34.36

high grade

3-2 1.73 0.1900

3-3 2.54 0.0880

2-3 1.73 0.0431

3-1 0.64 0.0630

2-1 0.54 0.0113

2-2 2.40 0.0211

1-2 1.59 0.0099

1-3 2.99 0.0127

Recovery, %

1-1 0.54 0.0130

Assay Head Calc Head Ro Conc Ro TailSample

4-1 0.81 0.0021

4-2 1.58 0.0021

4-3 3.01 0.0025

Cu, % Mo, % Cu, % Mo, % Cu, % Mo, % Cu, % Mo, % Cu, % Mo, % Cu MoBM-22 1.49 0.0108 146 39.54 0.244 30.72 0.184 0.030 0.0023 87.12 74.21BM-28 1.50 0.0089 146 43.43 0.196 35.16 0.160 0.030 0.0020 96.10 73.06 Additional collectorBM-32 1.53 0.0088 146 --- --- 35.67 0.160 0.040 0.0020 97.43 76.07 Timed concentratesBM-23 2.16 0.0267 167 48.28 0.590 39.86 0.445 0.030 0.0031 88.08 87.17BM-29 2.05 0.0221 167 52.32 0.510 43.01 0.417 0.030 0.0030 96.71 87.47 Additional collectorBM-33 2.15 0.0214 146 --- --- 43.39 0.371 0.030 0.0030 98.63 84.86 Timed concentratesBM-24 1.88 0.2006 138 45.45 5.230 39.16 4.180 0.030 0.0030 89.87 96.84BM-30 1.93 0.2034 138 47.37 5.040 37.29 3.903 0.020 0.0030 97.04 98.03 Additional collectorBM-25 1.72 0.0055 128 30.91 0.051 25.68 0.050 0.038 0.0024 96.35 49.33BM-31 1.63 0.0062 128 32.77 0.780 27.73 0.071 0.020 0.0020 98.33 61.59 Additional collector

(CA) CP-PY 3.40 0.0020 BM-46 3.63 0.0030 --- 32.33 0.010 28.82 0.005 0.030 0.0020 99.08 38.49*Micrometers

CA = Childs-Aldwinkle CP = ChalcopyriteVIX = Creek Breccia PY = Pyrite

CC = ChalcociteMo = Molybdenum

4-2(VIX) CP & PY

Recovery % Remarks

2.4

1.73

1.58

0.0211

0.1900

0.0021

GrindP80*

Ro Tail

1-2(CA) CP & BN 1.43 0.0099

2-2(CA) BN

3-2(CA) MO & CP

Assay HeadSample TestNo.

Calc Head Cl Conc Ro Conc


SGS FORM No. A263a-2/14/13 !


The effect of flotation retention time was also evaluated by MSRDI. The recovery - time profile for both copper and molybdenum from two composites are shown in Figure 13.1. Results indicate rapid flotation kinetics with over 95 percent of the copper recovered in three minutes with the flotation reagents used.

! Figure 13.1 - MSRDI Rougher Flotation Kinetics

After the batch test described above, MSRDI conducted locked cycle tests on the average grade composites. Results are summarized in Table 13.3. Concentrate grades between 32 and 62 percent copper were obtained with copper recoveries all above 95 percent. Molybdenum recoveries were proportional to the molybdenum head grade with the high-grade sample giving 94 percent recovery and the low-grade sample giving 28 percent recovery. As can be seen, flotation response of all the various mineralization types, as well as grades, was excellent averaging over 97 percent for copper and 72 percent for molybdenum.

Table 13.3 MSRDI Locked Cycle Test Results

Cum

ulat

ive

Rec

over

y (p

erce

nt)

50

63

75

88

100

Flotation Time (min)0 1 3 4 5

Test 32 Copper RecoveryTest 32 Moly RecoveryTest 33 Copper RecoveryTest 33 Moly Recovery


SGS FORM No. A263a-2/14/13 !

Test GrindCu, % Mo, % No. Cu, % Mo, % P80* Cu, % Mo, % Cu, % Mo, % Cu, % Mo, % Cu Mo

1-2 (CA) CP & BN 1.43 0.0099 BM-34 1.50 0.0120 146 44.04 0.260 34.36 0.208 0.050 0.0033 97.03 72.902-2 (CA) BN 2.40 0.0211 BM-35 2.14 0.0250 167 52.47 0.510 43.28 0.421 0.030 0.0042 98.66 82.343-2 (CA) MO-BN 1.73 0.1900 BM-36 1.62 0.1820 138 47.47 5.100 38.78 4.140 0.020 0.0087 99.04 94.831-2 (CA) 350-550 4.90 0.0037 BM-39 4.70 0.0051 116 61.83 0.037 53.35 0.034 0.250 0.0030 95.06 52.564-2 & 4-3 (VIX) CP & PY --- --- BM-40 2.37 0.0034 127 32.38 0.013 26.74 0.015 0.020 0.0030 99.37 28.151-3 (CA) 2.99 0.0120 BM-45 3.15 0.0150 --- 44.43 0.185 35.13 0.040 0.030 0.0020 99.13 86.122-3 (CA) BN 1.73 0.0430 BM-41 1.81 0.0320 --- 44.68 0.780 34.76 0.120 0.030 0.0010 98.53 97.053-3 (CA) MO & CC 2.54 0.0880 BM-44 2.44 0.0840 --- 43.27 0.146 33.34 0.240 0.090 0.0050 96.57 94.34(OR) 400-600 CC 1.42 0.0070 BM-42 1.65 0.0090 --- 45.98 0.210 31.52 0.030 0.170 0.0020 90.17 78.08(VIX) 670-710 CP & PY 1.28 0.0020 BM-46 1.96 0.0030 --- 40.06 0.020 30.25 0.010 0.020 0.0019 98.90 37.40

Average (arithmetic) 2.27 0.0419 2.33 0.0371 45.66 0.73 36.15 0.53 0.07 0.0034 97.25 72.38* Micrometers

CA = Childs-Aldwinkle CP = ChalcopyriteVIX = Creek Breccia PY = PyriteOR = Old Reliable BN = Bornite

CC = Chalcocite

Ro Conc Ro Tail Recovery, %Sample Assay Head Calc Head Cl Conc


In 2008, METCON conducted rougher flotation testing on composite samples identified as follows:

▪ Mid Grade Globe Breccia Composite ▪ High Grade Globe Breccia Composite ▪ Strongly Oxidized Copper Prince Composite ▪ Weakly to Unoxidized Copper Prince Composite ▪ Keel & American Eagle Composite

Head grades for the composites are shown in Table 13.4. It is noted that the molybdenum head grade in these composites is very low.

METCON conducted the rougher flotation tests on the composite samples at 38 percent and 22 percent solids and grind sizes of approximately 80 percent passing 175 microns, and 140 microns. The best results from the METCON tests are summarized in Table 13.5.

Table 13.4 METCON Composite Test Head Analysis

Sample IdentificationAssays

Cu (%)

Mo (%)

Fe (%)

Au (g/t)

Ag (g/t)

Mid Grade Globe Breccia Composite 0.87 0.003 3.84 N/A N/A

High Grade Globe Breccia Composite 1.43 0.003 4.04 N/A N/A

Strongly Oxidized Copper Prince Composite 2.87 0.005 2.99 0.03 3.40

Weakly to Unoxidized Copper Prince Zone Composite 2.35 0.010 3.14 0.05 5.00

Keel & American Eagle Composite 0.53 0.009 2.16 0.02 1.00

Table 13.5 METCON Rougher Flotation Test Results

Composite Sample ID Products

Wt. (%)

Assays ( % )

Distribution (%)

Cu (%)

Mo (%)

Cu (%)

Mo (%)

Mid Grade Globe Breccia Cu-Mo Ro. Concentrate 8.89 8.49 0.015 85.99 42.26

High Grade Globe Breccia Cu-Mo Ro. Concentrate 14.78 8.90 0.009 94.49 43.93

Strongly Oxidized Copper Prince Cu-Mo Ro. Concentrate 4.68 30.60 0.043 47.14 12.71

Weakly to Unoxidized Copper Prince Cu-Mo Ro. Concentrate 6.85 30.10 0.045 77.98 76.80

Keel & American Eagle Composite Cu-Mo Ro. Concentrate. 6.78 7.70 0.120 90.91 80.24


SGS FORM No. A263a-2/14/13 !


METCON results indicate that recovery is a function of oxidation and head grade. METCON conducted sulfidization tests with sodium hydrosulfide (NaHS) in an attempt to improve results from the oxidized composites. These tests did not result in significant recovery improvement. Redhawk geologists indicate that the amount of oxide material in the deposit is low.


SGS FORM No. A263a-2/14/13 !


13.1.2 Copper Molybdenum Separation Test Program

Limited testing was conducted by MSRDI to demonstrate the copper molybdenum separation process. The tests using NaHS, the conventional reagent for copper-molybdenum separation circuits, were not successful and additional testing to optimize reagents used in the bulk copper molybdenum flotation step and the subsequent copper molybdenum separation is recommended. It should be noted that it is difficult to undertake the proposed copper-molybdenum separation in the initial laboratory testing phase of the mining project due to lack of the appropriate quality and quantity of bulk copper-molybdenum concentrates.

MSRDI carried out the proposed copper-molybdenum separation testing on a composite sample made up of higher grade molybdenum samples from the Childs-Aldwinkle deposit (Samples 3-1, 3-2 and 3-3). These samples contained a mixture of chalcopyrite, bornite, and chalcocite, and had an average copper grade of 1.28 percent and an average molybdenum grade of 0.10 percent. In this testing program, ten kilograms of the composite sample were ground to the desired size distribution (P80 = 150 microns) and were subjected to flotation using the standard flotation conditions and reagents to obtain the final copper-molybdenum concentrates for the subsequent copper-molybdenum separation tests (Test No. BM-47). The bulk concentrate from the preceding test was split into eight equal portions of about 50 grams each and an additional sample for head assay. The eight samples were subjected to selective flotation techniques utilizing various copper or molybdenum depressants. The results of this testing program are summarized in the following Table 13.6.

The results of the scoping tests indicated that selective flotation of molybdenum from the bulk copper-moly concentrates produced using conventional copper depressants such as NaHS, ferro-cyanide, and hypochlorite was not effective. This lack of effectiveness may be attributed to the fact that the collectors S-5741 and Nalco 9743

Table 13.6 Summary Results of Copper-Moly Separation

Test No. Reagent Amount (lb/t) Remarks

BM-47-A NaHS 10.0 No depression of copper

BM-47-B Na-Ferrocyanide Up to 5.0 No depression of copper

BM-47-C Na-Cyanide Up to 5.0 No depression of copper

BM-47-D Na-Hypochlorite Up to 5.0 No depression of copper

BM-47-E Steaming 4 Hours No depression of copper

BM-47-F Oven-Drying 8 Hours Some depression of copper

BM-47-G CM-Guar Up to 3.0 Some depression of molybdenum


SGS FORM No. A263a-2/14/13 !


form very strong chemical bonds with the copper minerals. This strong bond renders typical copper depressants ineffective.

A single test evaluating the use of a milder xanthate collector to obtain a bulk copper-moly concentrate followed by selective flotation of molybdenum using NaHS reagent (Test BM-49) was run. Results of this test indicate that effective selective flotation of molybdenum from the bulk concentrate was achieved. The molybdenum grade of the rougher concentrate was 29.80 percent. This favorable result indicates that it should be possible to obtain acceptable molybdenum recovery at marketable grades using xanthate collectors but more testwork on representative samples is needed to confirm and optimize the process.

It should be noted that molybdenite occurrences in the Copper Creek deposits contain appreciable amounts of rhenium (up to 1100 ppm). Since the rhenium is contained in MoS2, crystals, it will be concentrated with molybdenum and can be recovered during the conversion of molybdenite concentrates to technical grade molybdic oxide.

In 2008, METCON also attempted a copper-molybdenum separation using NaHS. The test was run on a concentrate from the Keel & American Eagle composite sample. Unfortunately, the test head grade was very low, so the resulting test mass balance may not be representative. Consequently, it is not possible to verify the molybdenite separation process on this material. Results from the test are summarized in Table 13.7.

13.1.3 Bond Grinding Work Index

Two composite samples of Copper Creek mineralization (one comparatively soft and the other comparatively hard) were prepared and subjected to standard Bond Grinding Work Index tests at 100 mesh by MSRDI. The results of these tests revealed that the harder composite has a Work Index of 13.2 kWh/ton while the softer composite has a Work Index of 11.5 kWh/ton.

Table 13.7 METCON Copper – Molybdenum Separation Test Results

Products

Weight (%)

Assays ( % ) Distribution (%)

Cu (%)

Mo (%)

Fe (%)

Cu (%)

Mo (%)

Fe (%)

Molybdenum 3rd Cleaner Concentrate 0.01 12.54 22.02 11.16 0.32 38.18 0.07

Mo Rougher Tail (Cu Concentrate) 1.20 32.76 0.096 27.70 68.46 13.74 14.14

Calculated Head 0.57 0.008 2.35

Assay Head 0.53 0.009 2.16


SGS FORM No. A263a-2/14/13 !


13.1.4 Mineralogical Studies

The mineralogical studies indicated that the mineral grains were very coarse and that the mineralized material may be amenable to processing by heavy-media-separation techniques. For this reason, much of the work conducted at MSRDI was aimed at evaluating the response of the mineralization to a heavy-media-separation technique. The mineralogical studies and subsequent chemical analyses indicated the presence of tennantite, an arsenic-bearing mineral in some of the copper concentrates produced.

The general mineralogy distribution, as interpreted by AMT, indicates the following:

▪ Chalcocite occurs on the upper portions and down the margins of the breccia deposit. Significant chalcocite may exist to the 3,750 ft elevation. Chalcocite is noted to replace chalcopyrite and bornite. It is possible some of the bornite in this upper zone may be supergene in origin.

▪ Chalcopyrite forms the main hypogene zone. It occurs by itself and mixed with bornite or pyrite. It occurs mainly at the outer portions of the pipe and at depth.

▪ Bornite appears to occur in the core of the breccia body, usually with chalcopyrite. It does not seem to occur below 3,450 ft elevation.

▪ Below the 3,450 ft elevation, tennantite occurs, usually replacing chalcopyrite and bornite. Above 3,530 ft elevation the arsenic content is less then 10 ppm and does not exceed 1,000 ppm (0.10 percent) until the 3,460 ft elevation is reached.

MSRDI concluded the following from this phase of the testwork:

▪ The flotation response of all the mineralization types and grades to standard flotation operating conditions employed at the San Manuel concentrator is excellent.

In the single-stage flotation tests, copper recoveries ranged from plus 90 percent for lower grade composites to as high as 99 percent for higher grade mineralization, along with rougher concentrate grades ranging from about 22 percent copper to as high as 42 percent copper depending on the bornite content in the feeds.

▪ In the confirmatory locked-cycle flotation tests, the recoveries of copper in the final concentrates were 96 to 98 percent with grades varying from 27 to 42 percent copper depending on the bornite content in the feeds.


SGS FORM No. A263a-2/14/13 !


▪ It should be noted that the excellent results obtained in the above tests required a grind P80 equal to 140 microns.

▪ The results also showed that the recoveries of the by-product molybdenum were also relatively high and varied from a low of 26 percent to a high as 94 percent depending on the head grade of Mo in the composite samples.

▪ The Work Index determined for the composite samples investigated were as follows:

Softer Mineralization Composite: 11.5 kWh/ton Harder Mineralization Composite: 13.2 kWh/ton.

▪ In regard to the quality of the final flotation concentrates, the detailed chemical analysis for 22 elements including toxic elements indicated that all the final products were relatively clean of toxic elements except for unusually high arsenic contents in some of the concentrates.

▪ This presence of arsenic was attributed to tennantite occurrences in some localized areas of the Copper Creek deposits. It is important to acknowledge the presence of arsenic and to put in place appropriate steps need to keep the arsenic content below the smelter penalty limit in the final copper concentrates.

▪ Results of copper-molybdenum selective flotation tests on the bornite rich bulk flotation concentrates indicated that molybdenum depressants such as dextrine and starches were more effective in depressing molybdenite rather than sodium hydrosulfide, ferrocyanide and cyanide used for copper depression. Effective copper-molybdenum separations were also achieved by floating the molybdenum selectively with fuel oil prior to flotation of copper by conventional collectors. Similarly, effective separation of molybdenum from the copper-moly bulk concentrates was achieved by using the xanthate type collector for bulk concentration. In this case, the xanthate complex was decomposed by acid conditioning and then the copper was depressed with the conventional sodium hydrosulfide technique. Equally effective separation of molybdenum was achieved by this technique using the standard flotation reagents that were used by the San Manuel concentrator.

Review of the chemical analyses of the final copper concentrates from locked-cycle tests shown in Table 13.8 indicated that two out of eight composite samples showed high arsenic content. Since this high arsenic content could render the concentrates difficult to market or may involve a substantial penalty at the smelter. It is recommended that the AMT geological staff fully evaluate the extent of copper and arsenic mineralization (tennantite) in the Copper Creek Property. James Guthrie, AMT Chief Geologist indicated that the arsenic content is rather limited and that arsenic is not


SGS FORM No. A263a-2/14/13 !


present in any significant amount above the 3,460 ft elevation in the Copper Creek Property. Accordingly, the arsenic content in the final copper concentrates would be under the acceptable limits for a major portion of the mine life. If the arsenic content in the copper flotation concentrate becomes an issue, then blending with concentrates having a lower arsenic content will be utilized.


SGS FORM No. A263a-2/14/13 !


Table 13.8 Concentrate Analysis

Arsenic was assayed in the recent drilling completed by Redhawk. Results are summarized in Figure 13.2 support the AMT conclusion that the arsenic concentration is generally low but spotty and can be controlled by blending.

! Figure 13.2 - Redhawk 2007 Drill Results

Redhawk 2007 Drill Results Arsenic Concentration Vs. Drill Hole

0

1,000

2,000

3,000

4,000

5,000

6,000

7,000

CG

2C

G4

CG

5C

G6

CK3

2+0

CK3

5.5-

50C

P5G

10G

11G

13 G6

G7

G8

G9

G9A

NE-

6R

UM

-1S-

2 98

-1U

B-4R

VIX2

8-2

WC

-1LM

-10

LM-1

1LM

-9C

ATEC

HLM

-10

LM-9

RAE

-07-

019

RC

P-08

-026

RC

P-08

-027

RC

P-08

-028

RC

P-08

-030

RC

P-08

-038

RG

C-0

7-01

4R

GS-

08-0

41R

GS-

08-0

42R

MK-

07-0

15R

MM

-06-

001

RM

M-0

6-00

2R

MM

-06-

003

RM

M-0

7-00

4R

MM

-07-

007

RM

M-0

7-01

0R

MM

-07-

012

RM

M-0

7-01

3R

PE-0

8-03

3R

PE-0

8-03

4R

PE-0

8-03

5R

PE-0

8-03

6R

PE-0

8-03

7R

PE-0

8-03

9R

PE-0

8-04

0

Hole

Ars

enic

(ppm

)

Hole AverageHole Maximum


SGS FORM No. A263a-2/14/13 !

MSRDI Sample Cu Mo As Au Ag Hg Cd Pb Zn Sb Bi Co Mn Ni Se Te F MgO Cr Fe S ClNo. Identification % % % Oz/T Oz/T PPM % % % % % % % % % % PPM % % % % %

14985 BM-34 Cu Cl Conc 1-5 43.18 0.227 2.040 0.01 6.19 8.4 0.01 0.05 0.334 ND 0.15 0.01 0.06 0.038 0.03 ND ND 0.20 0.05 15.50 24.91 0.24Sample 1-2 (CA) CP & BN

14987 BM-35 Cu Cl Conc 1-5 52.52 0.494 4.250 0.11 3.56 0.6 0.01 0.02 0.294 ND 0.10 0.01 0.06 0.033 0.03 ND ND 0.30 0.05 9.55 23.14 0.14Sample 2-2 (CA) BN

14989 BM-36 Cu Cl Conc 1-5 48.43 0.520 0.020 0.08 5.12 0.6 0.01 0.01 0.308 ND 0.10 0.00 0.02 0.030 0.03 ND ND 0.17 0.04 13.04 25.90 0.11Sample 3-2 (CA) MO & BN

18752/3 BM-41 Cu Cl Conc 1-5 or 6 44.68 0.790 0.019 0.34 4.98 15.0 --- 0.06 0.128 ND 0.09 0.00 0.01 0.015 0.01 ND 0.001 0.15 0.02 4.54 26.96 0.10Sample 2-3 (CA) BN

18757/8 BM-42 Cu Cl Conc 1-5 or 6 45.98 0.200 0.009 0.01 1.28 8.0 --- 0.01 0.030 ND ND 0.05 0.01 0.034 0.01 ND 0.001 0.22 0.03 9.06 25.02 0.10Sample OR 400-600

19056/7 BM-44 Cu Cl Conc 1-5 or 6 43.27 1.460 ND 0.04 2.18 2.4 --- 0.04 0.006 ND 0.04 0.00 0.02 0.005 0.01 ND 0.003 0.02 0.02 16.31 9.63 0.09Sample 3-3 (CA) MO & CC

19058/9 BM-45 Cu Cl Conc 1-5 or 6 44.43 0.185 ND 0.08 3.78 1.6 --- 0.04 0.006 ND 0.06 0.00 0.00 0.007 0.00 ND 0.002 0.01 0.01 17.85 10.76 0.06Sample 1-3 (CA) CP & BN

19060/1 BM-45 Cu Cl Conc 1-5 or 6 40.06 0.022 ND 0.01 1.01 3.2 --- 0.02 0.070 ND ND 0.01 0.01 0.290 ND ND 0.002 0.02 0.05 26.21 12.60 0.06Sample VIX 670-710


13.1.5 Variability Second Cleaner Flotation Study on Variability Composite

In year 2012, METCON Research (METCON) conducted a preliminary second cleaner variability froth flotation study on 14 composite samples from Copper Creek Project. The composite samples were identified as follows:

▪ Composite 1 - Copper Grade in the 0.2 to 0.3 Percent Range ▪ Composite 2 - Chalcopyrite Dominant ≥ 0.2 to 0.5 Percent Copper ▪ Composite 3 - Chalcopyrite Dominant ≥ 0.5 Percent Copper ▪ Composite 4 - Bornite Moderate to Strong ≥ 0.2 to 0.5 Percent Copper ▪ Composite 5 - Bornite Moderate to Strong ≥ 0.5 Percent Copper ▪ Composite 6 - High Copper Grade ▪ Composite 7 - Mid Copper Grade ▪ Composite 8 - Low Copper Grade ▪ Composite 9 - SE Low Grade ▪ Composite 10 - SE Mod-High Grade ▪ Composite 11 - SE High Bornite ▪ Composite 12 - SW Low Copper Grade ▪ Composite 13 - SW Moderate High Copper Grade ▪ Composite 14 - SW High Bornite

Total copper, molybdenum, total gold and total silver assays for the above 14 composite samples are summarized in Table 13.9 below.

Table 13.9 Head Assays Summary of Results


Cu (%)

Mo (%)

Au (g/t)

Ag (g/t)

Composite 1 - Copper Grade in the 0.2 to 0.3 Percent Range 0.25 0.005 NA NA

Composite 2 - Chalcopyrite Dominant Copper Grade ≥ 0.2 to 0.5 Percent 0.27 0.002 NA NA

Composite 3 - Chalcopyrite Dominant, Copper Grade ≥ 0.5 Percent 0.91 0.030 NA NA

Composite 4 - Bornite Moderate to Strong, Copper Grade ≥ 0.2 to 0.5 Percent 0.28 0.011 NA NA

Composite 5 - Bornite Moderate to Strong, Copper Grade ≥ 0.5 Percent 0.82 0.012 NA NA

Composite 6 - High Copper Grade 0.85 0.007 NA NA

Composite 7 - Mid Copper Grade 0.37 0.006 NA NA

Composite 8 - Low Copper Grade 0.25 0.004 NA NA

Composite 9 - SE Low Copper Grade 0.30 0.001 0.01 1.30

Composite 10 - SE Moderate High Grade 0.61 0.005 0.01 2.30

Composite 11 - SE High Bornite 0.27 0.014 0.01 1.30

Composite 12 - SW Low Copper Grade 0.29 0.011 0.01 1.30

Composite 13 - SW Moderate High Copper Grade 0.62 0.002 0.02 1.60


SGS FORM No. A263a-2/14/13 !


Remarks: NA = Assays Not Available

The following comments relate to the chemical analyses conducted on the head samples.

▪ Total copper grade ranged from 0.25 percent to 0.91 percent. The highest copper grade of 0.91 percent is observed in Composite 3.

▪ Total molybdenum grade ranged from 0.001 percent to 0.232 percent. The highest molybdenum grade of 0.232 percent is observed in Composite 14.

▪ Gold and silver assays were conducted on Composites 9 through 14. Low gold and silver assays are observed on these composites.

Open cycle copper-molybdenum second cleaner flotation testing was conducted on 14 composite samples representing different zones from the Copper Creek Project. The grinding product size of P80 140 microns was used. The metallurgical data developed on the second cleaner flotation testing are summarized in Table 13.10 below.

Remarks: N/A, data not available

Composite 14 - SW High Bornite 0.63 0.232 0.05 3.10

Table 13.10 Cu-Mo Second Cleaner Flotation On Composite Samples

Sample ID

Cu-Mo Second Cleaner Concentrate Recovery (%)

Cu (%)

Mo (%)

Au (g/t)

Ag (g/t) Cu Mo Au Ag

Composite 1 - Copper Grade in the 0.2 to 0.3 Percent Range 28.80 0.56 1.20 NA 86.71 75.34 NA NA

Composite 2 - Chalcopyrite Dominant Copper Grade ≥ 0.2 to 0.5 Percent 30.50 0.39 1.40 NA 85.26 72.03 NA NA

Composite 3 - Chalcopyrite Dominant, Copper Grade ≥ 0.5 Percent 30.20 0.75 1.49 NA 87.23 73.76 NA NA

Composite 4 - Bornite Moderate to Strong, Copper Grade ≥ 0.2 to 0.5 Percent 41.80 2.28 3.95 NA 85.43 72.45 NA NA

Composite 5 - Bornite Moderate to Strong, Copper Grade ≥ 0.5 Percent 40.10 0.56 5.66 NA 77.17 80.67 NA NA

Composite 6 - High Copper Grade 31.10 0.20 0.96 NA 88.95 77.40 NA NA

Composite 7 - Mid Copper Grade 23.90 0.20 0.93 NA 87.36 66.46 NA NA

Composite 8 - Low Copper Grade 25.50 0.34 0.95 NA 82.78 65.97 NA NA

Composite 9 - SE Low Copper Grade 18.99 0.04 0.54 47 88.59 37.84 57.37 54.08

Composite 10 - SE Moderate High Grade 21.07 0.16 0.57 61 92.84 80.17 72.42 70.16

Composite 11 - SE High Bornite 21.84 1.07 0.41 56 88.27 87.33 45.47 53.87

Composite 12 - SW Low Copper Grade 20.84 0.79 0.73 46 85.07 86.69 57.74 49.11

Composite 13 - SW Moderate High Copper Grade 31.01 0.03 0.77 44 89.29 38.67 62.83 48.27

Composite 14 - SW High Bornite 31.50 12.30 3.59 154 91.81 97.06 82.40 78.80


SGS FORM No. A263a-2/14/13 !


The following comments relate to the open cycle copper-molybdenum second cleaner flotation tests on composite samples from Copper Creek Project.

▪ Copper recovery obtained ranged from approximately 77 percent to 93 percent. The highest copper recovery of approximately 93 percent was observed on Composite 10.

▪ The lowest copper recovery of approximately 77 percent was obtained on Composite 5.

▪ Molybdenum recovery achieved ranged from approximately 38 to 97 percent. Composite 14 provided the highest molybdenum recovery of approximately 97 percent.

▪ The lowest molybdenum recovery of approximately 38 percent was obtained on Composite 9.

▪ Complete mass balance for gold and silver was conducted on composites 9 through 14.

▪ Additional locked cycle flotation testing should be conducted to determine the ultimate concentrate and tailing values that will be realized when internal products are recycled as occurs in normal plant practice.

13.1.6 Copper-Molybdenum Separation on Two Composite Samples

In third quarter of year 2012, METCON conducted a copper-molybdenum separation on two composites identified as Open Pit and Underground from the Copper Creek Project. The composite samples were reconstituted per instructions received from Mr. Joe Sandberg, President and CEO of Redhawk Resources (Redhawk).

Approximately 700 kilograms of sample was prepared for the Open Pit Composite and approximately 1,100 kilograms of sample was prepared for the Underground Composite. Each composite sample was stage crushed to 100 percent minus 10 mesh and test charges of 1,000 grams were prepared for head assays, grind calibration and flotation testing. The test charges were saved in a freezer to minimize oxidation.

A test charge from each composite was selected at random, pulverized to 100 percent minus 150 mesh and a sample submitted for head assays. The results are summarized in Table 13.11 below.

Table 13.11 Head Assays on Composite Samples



SGS FORM No. A263a-2/14/13 !


The following comments relate to the chemical analyses conducted on the head samples.

▪ Total copper grade ranged from 0.33 percent to 0.41 percent. The highest copper grade of 0.41 percent is observed in the Underground Composite Sample.

▪ Both composite samples showed similar molybdenum head assay of 0.01 percent.

▪ Total silver grade ranged from 1.4 g/t to 3.5 g/t. The highest silver grade of 3.5 g/t is observed in the Underground Composite Sample.

▪ Low gold content is observed in both composite samples.

A copper/molybdenum concentrate was produced for each composite sample using the best rougher flotation and cleaner flotation conditions developed for each composite sample. The grinding product size of P80 140 microns was used for flotation circuit. Copper/molybdenum separation was conducted using sodium hydrosulfide and the molybdenum concentrate was subjected to at least five stages of cleaner flotation to produce the final molybdenum concentrate.

The following table summarizes the metallurgical data developed on the Cu-Mo concentrate production and the molybdenum fifth cleaner flotation conducted on the Cu-Mo concentrate produced from the Open Pit Composite.

Sample Identification Cu (%)

Mo (%)

Fe (%)

Ag (g/t)

Au (g/t)

ST (%)

Insol (%)

Open Pit Composite 0.33 0.01 2.82 1.4 0.01 1.33 84.86

Underground Composite 0.41 0.01 2.20 3.5 0.03 0.74 86.36

Table 13.12 Molybdenum Fifth Cleaner Flotation

Open Pit Composite Sample

Reagent

On Cu-Mo

Flotation

Flotation Products

Mass Recover

y (%)

Cumulative Grade Cumulative Recovery (%)

Cu (%)

Mo (%)

Fe (%)

Ag (g/t)

Au (g/t)

Total S

(%)Insol (%) Cu Mo Fe Ag Au Total

S Insol

S-5741 N9743

Mo 5th Cleaner

Concentrate

0.018 4.60 31.55 8.03 32 0.8

818.8

018.7

5 0.23 67.16 0.05 0.31 1.06 0.26 0.00

4

Cu Concentra

te1.353 22.5

5 0.02 26.85 45 0.5

626.5

015.0

086.5

5 2.60 12.23

33.52

51.91

27.76

0.242

Cu-Mo Concentra

te1.530 22.0

1 0.52 25.96 48 0.6

025.7

416.1

995.4

995.7

513.3

740.1

463.2

730.4

90.29

6


SGS FORM No. A263a-2/14/13 !


The following comments relate to molybdenum fifth cleaner flotation conducted on the Open Pit Composite on composite sample from Copper Creek Project.

▪ Copper and molybdenum recoveries of 95.49 percent and 95.75 percent respectively were obtained in the Cu-Mo bulk flotation using collectors S-5741 and N9743.

▪ A fifth molybdenum concentrate grade of 31.55 percent (67.16 percent molybdenum recovery) and a copper concentrate grade of 22.55 percent (86.55 percent copper recovery) were obtained.

▪ Copper and molybdenum recoveries of 95.48 percent and 93.64 percent respectively were obtained in the bulk flotation using collectors A-238 and MIBC as frother.


▪ Insoluble minerals reported to the fifth molybdenum concentrate and final copper concentrate are impacting grades and recovery. Further testing should be conducted to study insoluble minerals depression in order to increase molybdenum grade.

The following table summarizes the metallurgical data developed on the copper-molybdenum concentrate production and the molybdenum fifth cleaner flotation conducted on the copper-molybdenum concentrate produced from the Underground Composite.

A-238 MIBC

Mo 5th Cleaner 0.022 2.60 28.1

5 5.43 28 0.62

16.39

20.71 0.17 80.3

8 0.04 0.35 0.83 0.28 0.005

Cu Concentra

te1.251 23.7

0 0.01 25.80 46 0.7

424.0

015.0

587.7

3 1.98 11.05

33.04

58.10

23.96

0.222

Cu-Mo Concentra

te1.410 22.8

8 0.50 24.85 47 0.7

423.3

916.4

795.4

893.6

412.0

037.9

165.4

926.3

30.27

3

Table 13.13 Molybdenum Cleaner Flotation Kinetics

Underground Composite Sample

Flotation Products

Mass Recovery

(%)

Cumulative Grade Cumulative Recovery (%)

Cu (%)

Mo (%)

Fe (%)

Ag (g/t)

Au (g/t)

Total S

(%)Insol (%) Cu Mo Fe Ag Au Total

S Insol

Mo 5th Cleaner Concentrate 0.017 2.71 30.8

6 2.78 50 1.51 18.40 9.11 0.090 56.6

0 0.02 0.35 0.87 0.46 0.002

Cu Concentrate 1.422 30.45 0.02 26.80 78 1.55 23.7

5 8.88 83.09 3.69 15.79

43.57

73.40

49.27

0.148


SGS FORM No. A263a-2/14/13 !


The following comments relate to molybdenum fifth cleaner flotation conducted on the Underground Composite on composite sample from Copper Creek Project.

▪ Copper and molybdenum recoveries of 94.37 percent and 88.16 percent respectively were obtained in the copper-molybdenum bulk flotation.


▪ Insoluble minerals reported to the fifth molybdenum concentrate and final copper concentrate are impacting grades and recovery. Further testing should be conducted to study the insoluble minerals depression to increase molybdenum recovery and grade.

13.1.8 Comminution Testing

Drill core samples identified as RAE-08-45 and RMK-11-065 were received for comminution testing. Bond Crusher Work Index, Bond Rod Mill Work Index, Bond Ball Mill Work Index, Abrasion Index, JKTech Drop-weight and SMC Tests were conducted on the RAE-08-45 and RMK-11-065 drill core samples.

The Bond Crusher Work Index, Bond Rod Mill Work Index, Bond Ball Mill Work Index, Abrasion Index determinations were conducted at Phillips Enterprises, LLC (PE) in Colorado. The metallurgical data are summarized in the following paragraphs.

Bond Crusher Work Index determinations were conducted on the RAE-08-45 and RMK-11-065 samples by test protocol and the metallurgical data are summarized in Table 3.14 below.

Bond Rod Mill Work Index determinations were conducted by test protocol using 14-mesh (1,180 µm) closing screens. The metallurgical data are summarized in Table 3.15 below.

Cu-Mo Concentrate 1.660 29.64 0.50 26.0

0 81 1.50 23.49 9.25 94.37 88.1

617.8

852.8

383.0

556.8

80.18

1

Table 13.14 Bond Crusher Work Index

Sample ID CWi (kW-hr/st) CWi (kW-hr/mt)

RAE-08-45 6.92 7.63

RMK-11-065 8.77 9.67


SGS FORM No. A263a-2/14/13 !


Bond Ball Mill Work Index determinations were conducted by test protocol using 100-mesh (150 µm) closing screens. The metallurgical data are summarized in Table 3.16 below.

Abrasion index tests were conducted on a minus 3/4 inch/plus 1/2 inch fraction of each sample according to test protocol. The metallurgical data are summarized in Table 3.17 below.

Table 13.15 Bond Rod Mill Work Index

Sample ID RMWi (kW-hr/st) RMWi (kW-hr/mt)

RAE-08-45 12.88 14.2

RMK-11-065 11.97 13.2

Table 13.16 Bond Ball Mill Work Index

Sample ID BMWi (kW-hr/st) BMWi (kW-hr/mt)

RAE-08-45 14.22 15.73

RMK-11-065 13.62 15.01


SGS FORM No. A263a-2/14/13 !


The JKTech Drop-weight and SMC Tests were conducted on the RAE-08-45 and RMK-11-065 drill core samples at SGS Minerals Services in Lakefield, Canada (SGS). The JKTech drop-weight test was performed on the two samples. For the top fraction, the specimens were tested under the two possible orientations below, and the fragments were analyzed separately as follows:

▪ Vertical, i.e. the cylinder was sitting on its flat side and the hammer was hitting the other flat end.

▪ Horizontal, i.e. the cylinder was sitting on the round side and the hammer was hitting the other round side.

The data was interpreted by Contract Support Services (CSS), the North American agent for JKTech. The overall test results are summarized in Table 3.18. The effect of orientation on the overall interpretation and parameters determination was very marginal, so the combined results were used.

Both samples were characterized as hard with respect to resistance to impact breakage (A x b). The sample RMK-11-065 and sample RAE-08-45 were classified with respect to resistance to abrasion breakage (ta) as very hard and hard respectively.

13.2 Process Flowsheet

The simplified process flowsheet (Figure 13.3) was developed for processing 25,000 tpd with an overall availability of 92.5 percent. The basis for the flowsheet, capital cost and operating cost are given in Table 13.19.

Table 13.17 Abrasion Index

Sample ID Ai

RAE -08-45 0.2479

RMK-11-065 0.3112

Table 13.18 JKTech Drop-weight and SMC Test Results

Sample ID A b A x b Percentile of Hardness ta

Percentile of Hardness

Density (g/cm3)

RMK-11-065 88.3 0.34 30.0 87 0.23 91 2.63

RAE-08-45 100.0 0.30 30.0 87 0.27 85 2.55


SGS FORM No. A263a-2/14/13 !


Table 13.19 Process Plant Design Parameters

Item Parameter

Short Tons per day 25,000

Availability (%) 92.50

Feed Grade

% Cu 0.77

% Mo 0.015

oz/ton Ag 0.056

Cu Concentrate grade

% Cu 30.0

Mo Concentrate grade

% Mo 47.0

Recovery %

Cu 92.0

Mo 78.0

Ag 50.0


SGS FORM No. A263a-2/14/13 !


!

Figure 13.3 - Redhawk Summary Flowsheet


SGS FORM No. A263a-2/14/13 !


14.0 MINERAL RESOURCE ESTIMATES

14.1 Summary

The December 2012 mineral resource is a combination of the re-statement of the October 2008 mineral resource for the five breccia deposits and the re-estimation of the Keel - American Eagle porphyry deposit. No new drilling has been done in the breccias so the October 2008 resource models are used with the resources updated to reflect a new calculation of the copper equivalent. The Old Reliable is reported separate from the other breccia deposits because it is anticipated to be mined by an open pit method, while the remaining breccia deposits would be mined underground. The Keel - American Eagle deposit has been re-estimated to include the additional drilling results since March 2012 (data cutoff date for the June 2012 resource) and to model the deposit for a potential underground mining approach.

Grades in the updated Keel - American Eagle resource model were estimated using inverse distance to the fourth power (ID4) in three estimation domains defined by Redhawk geologists as copper grade shells based on 20ft drill hole composites. Zone 5 was defined by composite grades where total copper greater than or equal to (>=) 0.50 percent, Zone 4 was defined by composite grades where total copper >= 0.40 percent and less than (<) 0.50 percent, and Zone 3 was defined by composite grades where total copper >=0.30 and < 0.40 percent. All three zones were estimated with a 350 ft by 350 ft (horizontal) and 275 ft (vertical) search. Blocks in Zone 3 were estimated using composites in Zones 3 and 4, blocks in Zone 4 were estimated by using composites in Zones 3, 4, and 5, and blocks in Zone 5 were estimated using composites in Zones 4 and 5.

The five breccia models used the outlines of the breccia bodies as limits to the grade estimates. There is isolated mineralization to the exterior of the breccia bodies and the estimate of grade of this mineralization was made using a polygonal estimation with a limited search distance. The Childs-Aldwinkle and Mammoth grade models were estimated using indicator kriging and the Old Reliable, Globe and Copper Prince grade models used ordinary kriging. A summary of the development of these resource models is presented later in this section and the details are found in the October 2008 technical report.

Tonnages were calculated assuming a constant density of 12.5 cu ft/ton (2.56 g/cc). Resources were classified as measured, indicated or inferred based on the number of holes within the search and the kriging variance. In preparing this estimate IMC has assumed that all of the data supplied to it, including assay and survey data, are correct to within normally-accepted limits of error.

The mineral resource in Table 14.1 is tabulated at different copper equivalent (CuEq) cutoffs for each of three areas:


SGS FORM No. A263a-2/14/13 !


▪ The near surface Old Reliable breccia is tabulated using a 0.40 percent CuEq cutoff to reflect a potential open pit resource.

▪ The breccia deposits of Globe, Copper Prince, Childs-Aldwinkle and Mammoth are tabulated using a 0.75 percent CuEq cutoff to reflect a selective, underground mining approach.

▪ The Keel - American Eagle porphyry is tabulated at a 0.50 percent CuEq cutoff to reflect a bulk underground mining approach.

The copper equivalent grade is based on the metal prices and estimate of recoveries for copper, molybdenum, silver and gold (estimated in the breccia deposits only).

The copper equivalent calculations are:

Keel - American Eagle: CuEq% = Cu% + 3.875 x Mo% + 0.3636 x Ag (opt)

Breccia Deposits: CuEq% = Cu% + 3.875 x Mo% + 0.3636 x Ag (opt) + 8.89 x Au (opt)

Metal Price Recovery

Copper $2.75/lb 90%

Molybdenum $12.00/lb 80%

Silver $20.00/oz 90%

Gold $1100/oz 40%


SGS FORM No. A263a-2/14/13 !


14.2 Drilling and Assaying

Table 14.2 summarizes basic drilling and copper-molybdenum assaying statistics for the assay data base supplied to IMC. (Note that the sum of the holes in the individual categories exceeds the total number of holes by 56 because 56 holes were drilled using more than one drilling method.)

Table 14.1 December 2012 Mineral Resource Estimate

Deposit & CuEq Cutoff

Class Ktons Copper, %

Moly, %

Silver, opt Gold, opt CuEq, %

Old Reliable Breccia 0.40%

Measured 2,727 0.77 0.011 0.081 - 0.87

Indicated 255 0.60 0.008 0.034 - 0.65

M&I 2,982 0.75 0.011 0.077 - 0.85

Inferred 79 0.68 0.007 0.013 - 0.72

Breccia Deposits 0.75%

Measured 4,462 1.52 0.013 0.142 0.002 1.64

Indicated 1,272 1.86 0.035 0.111 0.003 2.07

M&I 5,734 1.60 0.018 0.135 0.002 1.74

Inferred 774 2.09 0.038 0.082 0.003 2.29

Keel – Am. Eagle 0.50%

Measured 29,765 0.79 0.020 0.100 - 0.90

Indicated 115,218 0.70 0.012 0.030 - 0.76

M&I 144,983 0.72 0.013 0.050 - 0.79

Inferred 85,841 0.68 0.014 0.040 - 0.75

Table 14.2 Drilling and Assaying Statistics, All Data

Diamond Core

Reverse Circulatio

nRotary/

PercussionDrifts/ Raises

Not Classifie

d Total

Number of holes 370 75 95 56 11 607

Total drilling, ft 524,818 35,787 89,131 3,056 6,730 659,521

Assayed for copper, ft 420,437 29,748 57,084 2,878 5,660 515,807




Assayed for molybdenum, ft 318,690 17,758 43,477 110 4,700 384,734




SGS FORM No. A263a-2/14/13 !


Table 14.3 summarizes drilling by the mineralized areas within the block model used for the Keel - American Eagle mineral resource update. This is a sub-set of the total drill hole data base which is presented in Section 10.



SGS FORM No. A263a-2/14/13 !


Drill hole locations at Copper Creek are shown in Figure 14.1. Numbered areas 4 through 6 illustrate the limits of the Childs-Aldwinkle, Mammoth, and American Eagle areas as defined in the block model used to update the Keel – American Eagle mineral resources.

!

Figure 14.1 – Drill Hole Locations

Holes are mostly vertical in the American Eagle area and spaced a few hundred feet apart on average. The holes that test the higher-grade breccias to the northwest are mostly inclined and hole spacing is highly variable. The drilling in the areas of Globe, Copper Prince, and Old Reliable were not included in the current resource model since these areas have no new data since the 2008 mineral resource estimate. Most of the deeper holes have down hole surveys that generally show deviations of less than 5

Table 14.3 Drilling Statistics by Mineralized Area

Mineralized Area

Number of

Holes

Total Drilling,

ftNumber of Intervals

Intervals Assayed

for Copper

Intervals Assayed for Moly

Intervals Assayed for Silver

Childs-Aldwinkle 58 45,634 2,501 2,380 2,042 656

Mammoth-Keel 82 138,556 11,884 11,666 11,472 6,224

American Eagle 92 205,243 19,974 19,783 15,754 10,352


SGS FORM No. A263a-2/14/13 !

N6

5

4


degrees. Hole diameters and recoveries were not recorded in the data base supplied to IMC.

For the Copper Creek Keel - American Eagle model, three copper grade shell zone boundaries were interpreted by Redhawk Copper geologists on each 20ft level. These boundaries were total copper >= 0.50 percent (Zone 5), total copper = >0.40 percent, < 0.50 percent (Zone 4), and total copper >=0.30 percent, < 0.40 percent (Zone 3). These outlines were used to tag the blocks in the block model. All other blocks were in Zone 0. Composite data was “dipped” into the block model and the composites were assigned a zone code based on which grade shell they were located. Table 14.4 summarizes the drilling by the grade shell domain.

14.3 Compositing, Grade Statistics

Assays were composited into 20 ft bench composites to match the 20 ft level height in the resource model. Copper and molybdenum assays and composite statistics are compared in Table 14.5. Table 14.6 summarizes 20 ft level composite statistics by defined grade zone.

Mean composite grades are lower than mean assay grades because assay intervals tend to be shorter in higher-grade zones.

Table 14.4 Composite Statistics by Zone Boundaries

Grade ShellTotal Drilling,

ftNumber of Intervals


Copper

Intervals Assayed for Molybdenum


Silver

Zone 5 28,003 1,385 1,384 1,298 724

Zone 4 12,890 638 637 566 369

Zone 3 23,010 1,144 1,138 1,010 665

Zone 0 333,850 16,061 12,822 10,907 6,437

Outside Model 263,763 13,027 9,354 5,120 2,552

Total 659,517 32,255 25,335 18,901 10,747

Table 14.5 Assay and 20ft Bench Composite Statistics, All Data

Copper, % Molybdenum, %

Number Mean St. Dev. Max. Number Mean St. Dev. Max

Assays 56,099 0.321 0.774 23.70 41,570 0.0058 0.031 2.15

Composites 25,335 0.290 0.573 9.55 18,901 0.0055 0.022 1.19


SGS FORM No. A263a-2/14/13 !


Figure 14.2 shows log-transformed cumulative frequency plots of 20 ft composite copper and molybdenum grades for all Copper Creek samples. Both distributions are effectively lognormal with no indication of high-grade outlier populations.

! Figure 14.2 - 20ft Composite Cumulative Frequency, Copper and Molybdenum, All Data

Table 14.6 20ft Bench Composites Statistics by Grade Zones

Grade ZoneCopper, % Molybdenum, %

Number

Mean St. Dev. Max. Number

Mean St. Dev. Max.

Zone 5 1,384 0.818 0.497 4.81 1,298 0.019 0.034 0.300

Zone 4 637 0.401 0.066 1.15 566 0.009 0.013 0.091

Zone 3 1,138 0.313 0.046 0.459 1,010 0.007 0.009 0.080

Zone 0 12,822 0.242 0.501 8.75 10,907 0.004 0.024 1.186


SGS FORM No. A263a-2/14/13 !


Figure 14.3 shows log-transformed cumulative frequency plots of 20 ft composite copper grades by grade shell zone. The Zone 0 (No Zone on the plot) includes drill hole composites within the Mammoth and Childs-Aldwinkle breccias which were not modeled by the grade shells. The higher grade composites within these breccias account for the kink in the cumulative frequency curve for this zone. This is also present in Figure 14.4 for molybdenum, although more gradual.

!

Figure 14.3 - 20ft Composite Cumulative Frequency, Copper, by Zone

Figure 14.4 shows log-transformed cumulative frequency plots of 20 ft composite molybdenum grades by zone. Distributions are lognormal in all zones.


SGS FORM No. A263a-2/14/13 !


! Figure 14.4 - 20ft Composite Cumulative Frequency, Molybdenum, by Zone

Reviews failed to detect any very-high-grade “outlier” copper or molybdenum composites, so no cap grades were applied before grades in the model were estimated.

14.4 Variograms, Distribution of Mineralization

Copper variograms at Copper Creek were looked at separately in all the defined grade zones. Reasonable results were obtained with covariance variograms run on 20 ft copper composites within each of these zones. A combined zone variogram is shown in Figure 14.5. The results of the variograms are summarized in Table 14.7.


SGS FORM No. A263a-2/14/13 !


These ranges are restricted in the horizontal distances by the size of the zone boundaries. Zone 3 has the longest vertical range. In Zones 4 and 5 the vertical range is shorter than the horizontal range.

Molybdenum variograms were not constructed because molybdenum has a lesser economic contribution to the overall copper equivalent value and the molybdenum grade tends to track copper grade.

Table 14.7 Covariance Copper Variograms, 20ft Composites

Range, ft Nugget Sill

All Zones - All horizontal directions 280 0.06 0.166

All Zones - Vertical 330 0.08 0.166

All Zones - Omnidirectional 290 0.06 0.166

Zone 3 – All horizontal 356 0.001 0.002

Zone 3 – Vertical 372 0.00 0.002

Zone 3 – Omnidirectional 341 0.001 0.002


Zone 4 - Vertical 262 0.002 0.004

Zone 4 - Omnidirectional 294 0.002 0.004


Zone 5 - Vertical 278 0.13 0.247

Zone 5 - Omnidirectional 222 0.11 0.247


SGS FORM No. A263a-2/14/13 !


! Figure 14.5 - Omnidirectional Covariance Copper Variogram, 20ft Copper Composites

All Zones


SGS FORM No. A263a-2/14/13 !


The distribution of copper mineralization relative to drilling at Copper Creek is illustrated in Figures 14.6 and 14.7, which show 20 ft composite copper grades on EW section 635500N in the American Eagle area and EW section 636500N through the Mammoth-Keel/American Eagle area (both section looking north).

! Figure 14.6 - 20ft Composite Copper Grade, Section 635500N, American Eagle

0-0.3% Blue, 0.3-0.4% Green, 0.4-0.5% Orange, >0.5% Pink, Black=Un-assayed


SGS FORM No. A263a-2/14/13 !


! Figure 14.7 - 20ft Composite Copper Grade, Section 636500N, Mammoth-Keel & American Eagle

(Mammoth-Keel on left, American Eagle on right) 0-0.3% Blue, 0.3-0.4% Green, 0.4-0.5% Orange, >0.5% Pink, Black=Un-assayed


SGS FORM No. A263a-2/14/13 !


Drilling at American Eagle is reasonably evenly-spaced, and on the Figure 14.6 section it defines a broadly continuous mineralized zone at depth which is open in some directions. Drilling outside of the higher grade intersections has intervals which were not assayed. Redhawk continues to add to the drill hole data base with new drilling and the assaying of intervals in previous drilled holes which were not assayed. This is an ongoing program to make the data base more complete.

14.5 Grade Estimation

The Copper Creek underground model extends from 63,3860N to 63,8400N, from 40,1000E to 40,800E and from -100 to 3800 ft elevation. With a model block size of 20x20x20ft it contains 350 columns, 227 rows and 195 tiers for a total of 15,492,750 blocks. This model was developed to update the mineral resource for the Keel - American Eagle porphyry. The 20 x 20 x 20 ft block was selected as a reasonable mining unit for a room and pillar underground mining approach which is being considered for the porphyry zones. The block model extends into the breccia areas, but mineral resources for these areas are tabulated for the 2008 block models and thus were not modeled even though they fall within this block model framework.

14.5.1 Selection of Grade Estimation Operator

IMC's experience in porphyry copper deposits where mined-model comparisons are available shows that a higher-power inverse distance operator is usually needed to match the model grade distribution to the blast hole grade distribution, and that the blast hole grade distribution is usually best matched when this operator “splits the difference” between ordinary kriging (OK), which tends to overstate tons and understates grade, and nearest-neighbor polygons (NNP), which tends to understate tons and overstate grade.

Using these criteria it was found that an inverse distance to the fourth power (ID4) operator was appropriate for Copper Creek. Figure 14.8 compares the copper block grade distributions obtained using this operator with the distributions obtained using ordinary kriging (OK) and nearest-neighbor polygons (NNP) in the Copper Creek model for initial block model runs.


SGS FORM No. A263a-2/14/13 !


!

Figure 14.8 - Model Block Grade Distributions: OK (Green), NNP (Blue), ID4 (Red)

14.5.2 Internal Boundaries and Grade Estimation Searches

Grade shell zone boundaries were interpreted and digitized by the Redhawk geologists. This information was assigned to the block model and to the composite data. Figure 14.9 shows the 1980 bench with the grade zone boundaries and the 20 ft composites for that bench. Zone 3 is total copper >=0.30 percent and < 0.40 percent, Zone 4 is total copper >=0.40 percent and < 0.50 percent and Zone 5 is total copper >= 0.50 percent.


SGS FORM No. A263a-2/14/13 !


!

Figure 14.9 - Grade Zone Boundaries with Composites on the 1980 Bench Zone 3 is Blue, Zone 4 is Green, and Zone 5 is Red.

Copper, molybdenum and silver were estimated using a 350 ft by 350 ft horizontal and 275 ft vertical search and a maximum/minimum of 12/2 20 ft composites in all three zones. The number of composites from a single drill hole was restricted to 6 composites. Blocks in Zone 3 were estimated with composites in Zones 3 and 4. Blocks in Zone 4 were estimated with composites in Zones 3, 4, and 5. Blocks in Zone 5 were estimated with composites in Zones 4 and 5. Blocks in Zone 0 were not estimated. This estimation did not assign grades to all the blocks within the grade shells. A second estimation pass was made in Zones 3, 4, 5 with the search radius expanded to 800 ft by 800 ft horizontal and 275 ft vertical to fill in more of the blocks within the shell zones, but the results of this pass were not included in the mineral resource estimate.

14.5.3 Treatment of Un-assayed Intervals

Within the grade shells used to estimate the porphyry mineral resource there are a total of 3,167 composites of which 3,159 have values for copper (99.7%), 2,874 has values for molybdenum (90.7%) and 1,758 have values for silver (55.5%). Copper is the largest contributor to the copper equivalent calculation, followed by molybdenum and silver making a minor contribution. The copper grades are complete throughout the grade shell zones and molybdenum has coverage of 90 percent. No default values have been assigned to either the assay or composite data bases.


SGS FORM No. A263a-2/14/13 !


14.5.4 Model Results

In the zone assigned blocks 9 percent of the composite intervals that are assayed for copper are not assayed for molybdenum, 841,594 blocks in the IMC model are assigned copper grades, 838,024 blocks are assigned molybdenum grades and 213,683 blocks are assigned silver grades. At a 0.5 percent copper cutoff, however, almost all of the blocks that are assigned copper grades are assigned molybdenum grades (318,882 vs. 316,128). Silver has fewer assigned at 122,388 blocks.

Block grade distributions in the IMC model are summarized in the following figures.

▪ Figure 14.10: Block copper grades along the 635500N drill hole composite grade section shown in Figure 14-6.

▪ Figure 14.11: Block copper grades along the 636500N drill hole composite grade section shown in Figure 14.7.

▪ Figure 14.12: Block copper grades on the 1980 bench showing zone boundaries, drill hole composite grade bench shown in Figure 14.9.

▪ Figure 14.13: Plan map showing grade-thickness of copper in model blocks

▪ Figure 14.14: Plan map showing grade-thickness of molybdenum in model blocks

These figures include measured, indicated and inferred blocks.


SGS FORM No. A263a-2/14/13 !


! Figure 14.10 - Copper Block Grade, East-West Section at 635000N, American Eagle

0-0.3% Blue, 0.3-0.4% Green, 0.4-0.5% Orange, >0.5% Pink, Blank=Un-estimated


SGS FORM No. A263a-2/14/13 !


! Figure 14.11 - Copper Block Grade, East-West Section at 636500N, Keel - American-Eagle



SGS FORM No. A263a-2/14/13 !


! Figure 14.12 - Copper Block Grades, Bench 1980



SGS FORM No. A263a-2/14/13 !

N


! Figure 14.13 - Grade-Thickness Copper in Model Blocks, Feet x Percent, Copper > 0.5%

100-200 Blue, 200-500 Green, 500-1,000 Orange, >1,000 Pink (Map Grid is 1000 x 1000 ft)


SGS FORM No. A263a-2/14/13 !

N


! Figure 14.14 - Grade-Thickness Molybdenum in Model Blocks, Feet x Percent, Copper > 0.5%

5-20 Blue, 20-50 Green, 50-100 Orange, >100 Pink (Map Grid is 1000 x 1000 ft)


SGS FORM No. A263a-2/14/13 !

N


14.5.5 Resource Classification

The classification of the mineral resource for the Keel – American Eagle porphyry started with all the resource being classified as inferred if the copper grade was estimated for a block. The confidence was increased as further restrictions were applied based on the number of composites or drill holes used to estimate a block’s grades.

Resources were initially classified as indicated where composites with copper values from three or more holes were present, within the search ellipse and the number of composites within the search ellipse was twelve or greater. Additionally for Zone 3, a block was classified as indicated if there were five or more holes present, within the search ellipse. For Zones 4 and 5 a block was classified as indicated if there were four or more holes present, within the search ellipse.

The three-hole-minimum criterion for the zones is supported by kriging variances, which are a measure of the errors to which individual block grade estimates are subject. Figure 14.15 plots kriging variance against the number of drill holes in the search for each zone, these show kriging variance increasing gradually as the number of holes within the search decreases from twelve to three but increasing more rapidly as the number falls below three. This inflection defines the three-hole minimum as an appropriate confidence threshold for segregating inferred from indicated material.

! Figure 14.15 Kriging Variance Vs. Number of Holes in Search Ellipsoid


SGS FORM No. A263a-2/14/13 !


Figure 14.15 (Continued)

!

!

0.0008

0.0013

0.0018

0.0023

0.0028

1 2 3 4 5 6 7 8 9 10 11 12

Copper Creek Zone 4

kvar

0.06

0.08

0.10

0.12

0.14

0.16

0.18

1 2 3 4 5 6 7 8 9 10 11 12

Copper Creek Zone 5

kvar


SGS FORM No. A263a-2/14/13 !


The indicated classification was sub-divided into measured and indicated by using the number of drill holes within a shorter distance from the block center. To be classified as measured there either had to be six or more drill holes used to estimate a block or there had to be 3 drill holes within 200 ft of the block.

The final resource classifications are strictly applicable only to copper grades, but since molybdenum and silver make only a comparatively small economic contribution the classifications have also been applied to these grades as well.

Resource classification results are summarized in Figures 14.13 and 14.14, which show indicated and inferred blocks along the east-west sections at 635500N and 636500N drill hole composite grade sections used in Figures 14.5, 14.6, 14.8 and 14.9 and in the copper grade-thickness plan maps above a 0.5 percent cutoff shown in Figure 14.15 (measured plus indicated) and Figure 14.16 (inferred).


SGS FORM No. A263a-2/14/13 !


! Figure 14.16 - Resource Classification, Section 635500N, American Eagle

Blue = Inferred, Green = Indicated, Red = Measured


SGS FORM No. A263a-2/14/13 !


! Figure 14.17 - Resource Classification, Section 636500N, Mammoth-Keel and American Eagle

Blue = Inferred, Green = Indicated, Red = Measured


SGS FORM No. A263a-2/14/13 !


! Figure 14.18 - Grade-Thickness Copper, Feet-Percent, Measured & Indicated Blocks, Copper >

0.5% 100-200 Blue, 200-500 Green, 500-1,000 Orange, >1,000 Pink (Map Grid is 1000 x 1000 ft)


SGS FORM No. A263a-2/14/13 !

N


! Figure 14.19 - Grade-Thickness Copper, Feet-Percent, Inferred Blocks, Copper > 0.5%

100-200 Blue, 200-500 Green, 500-1,000 Orange, >1,000 Pink (Map Grid is 1000 x 1000 ft)


SGS FORM No. A263a-2/14/13 !

N


14.6 2008 Breccia Resource Models

The mineral resources are summarized here and described in detail in the Independent Mining Consultants, Inc. (IMC) report titled: “Copper Creek 2008 Mineral Resource, Pinal County, Arizona, USA, Technical Report”, dated October 28, 2008 and filed on SEDAR on October 29, 2008. The five breccia deposits (Copper Prince, Globe, Old Reliable, Childs-Aldwinkle and Mammoth) contain copper mineralization with associated molybdenum (moly), gold and silver, in potentially economic concentrations.

The drilling used for the mineral resource estimates (the footage is only what is within each model area and does not include the total footage of drilling from the surface) is:

▪ Mammoth Breccia 63 surface drill holes totaling 95,745 ft

▪ Childs-Aldwinkle Breccia 85 surface holes and 6 underground holes totaling 90,064 ft

▪ Old Reliable Breccia 49 surface holes, 61 underground holes, 31 channel samples totaling 41,913 ft

▪ Copper Prince 18 surface holes totaling 8,438 ft

▪ Globe 23 surface holes totaling 16,054.8 ft

The five breccia models used the outlines of the breccia bodies as limits to the grade estimates. There is isolated mineralization to the exterior of the breccia bodies and the estimate of grade of this mineralization was made using a polygonal estimation with a limited search distance. The Childs and Mammoth grade models were estimated using indicator kriging and the Old Reliable, Globe and Copper Prince grade models used ordinary kriging. Block size in the breccias is 15’ x 15’ x 15’ based upon the assumption of the potential for blast-hole cut and fill mining methods.

14.7 Mineral Resource

Tonnages were calculated assuming a constant density of 12.5 cu ft/ton (2.56 g/cc). Resources were classified as measured, indicated or inferred based on the number of holes within the search ellipse. In preparing this estimate IMC has assumed that all of the data supplied to it, including assay and survey data, are correct to within normally-accepted limits of error.

The mineral resource in Table 14.8 is tabulated at different copper equivalent (CuEq) cutoffs for each of three areas:


SGS FORM No. A263a-2/14/13 !


▪ The near surface Old Reliable breccia is tabulated using a 0.40 percent CuEq cutoff to reflect a potential open pit resource. This resource has been confined within a pit shell based on process plus G&A costs of $10.00/ton and mining costs of $2.00/ton.

▪ The breccia deposits of Globe, Copper Prince, Childs-Aldwinkle and Mammoth are tabulated using a 0.75 percent CuEq cutoff to reflect a selective, underground mining approach.

▪ The Keel - American Eagle porphyry is tabulated at a 0.50 percent CuEq cutoff to reflect a bulk underground mining approach.

The copper equivalent grade is based on the metal prices and estimate of recoveries for copper, molybdenum, silver and gold (estimated in the breccia deposits only).

The copper equivalent calculations are:

Keel – American Eagle: CuEq% = Cu% + 3.875 x Mo% + 0.3636 x Ag (opt) Breccia Deposits: CuEq% = Cu% + 3.875 x Mo% + 0.3636 x Ag (opt) + 8.89 x Au (opt)

Metal Price Recovery

Copper $2.75/lb 90%

Molybdenum $12.00/lb 80%

Silver $20.00/oz 90%

Gold $1100/oz 40%


SGS FORM No. A263a-2/14/13 !


Table 14.9 is a summary of the Breccia deposits by the individual deposits at the 0.75 percent CuEq cutoff. Table 14.10 shows the Keel - American Eagle porphyry at different cutoff grades to show the distribution of mineralization with the mineral resource at 0.50 percent CuEq cutoff highlighted.

14.8 December 2012 Mineral Resource Estimate

Deposit & CuEq Cutoff Class Ktons

Copper, % Moly, % Silver, opt Gold, opt CuEq, %

Old Reliable Breccia 0.40%

Measured 2,727 0.77 0.011 0.081 - 0.87

Indicated 255 0.60 0.008 0.034 - 0.65

M&I 2,982 0.75 0.011 0.077 - 0.85

Inferred 79 0.68 0.007 0.013 - 0.72

Breccia Deposits 0.75%

Measured 4,462 1.52 0.013 0.142 0.002 1.64

Indicated 1,272 1.86 0.035 0.111 0.003 2.07

M&I 5,734 1.60 0.018 0.135 0.002 1.74

Inferred 774 2.09 0.038 0.082 0.003 2.29

Keel – Am. Eagle 0.50%

Measured 29,765 0.79 0.020 0.100 - 0.90

Indicated 115,218 0.70 0.012 0.030 - 0.76

M&I 144,983 0.72 0.013 0.050 - 0.79

Inferred 85,841 0.68 0.014 0.040 - 0.75


SGS FORM No. A263a-2/14/13 !


Table 14.9 Individual Breccia Deposits at 0.75% CuEq Cutoff

!

Breccia CuEq Cutoff ktons CuEq, % Cu, % Mo, % Ag, opt Au, opt

Childs 0.75 830 2.26 1.88 0.067 0.188 0.006Prince 0.75 253 2.07 2.02 0.003 0.086 0.001Globe 0.75 38 1.20 1.17 0.004 0.032 0.000

Mammoth 0.75 3,341 1.46 1.40 0.001 0.136 0.001Total 0.75 4,462 1.64 1.52 0.013 0.142 0.002

ktons CuEq, % Cu, % Mo, % Ag, opt Au, opt


Mammoth 0.75 242 2.78 2.71 0.001 0.161 0.001Total 0.75 1,272 2.07 1.86 0.035 0.111 0.003

ktons CuEq, % Cu, % Mo, % Ag, opt Au, opt

Childs 0.75 1,560 2.14 1.79 0.063 0.153 0.005Prince 0.75 450 2.00 1.94 0.005 0.079 0.001Globe 0.75 141 1.13 1.10 0.003 0.051 0.001

Mammoth 0.75 3,583 1.55 1.49 0.001 0.138 0.001Total 0.75 5,734 1.74 1.60 0.018 0.135 0.002

Breccia CuEq Cutoff ktons CuEq, % Cu, % Mo, % Ag, opt Au, opt


Mammoth 0.75 49 2.38 2.34 0.002 0.101 0.000Total 0.75 774 2.29 2.09 0.038 0.082 0.003

Measured

Indicated

Measured + Indicated

Inferred


SGS FORM No. A263a-2/14/13 !


Table 14.10 Keel - American Eagle Mineral Resource by Cutoff Grade

!

CuEqCutoff ktons CuEq, % Cu, % Mo, % Ag, opt

1.00 7,893 1.44 1.24 0.038 0.170.75 15,179 1.16 1.00 0.028 0.140.60 24,873 0.97 0.85 0.021 0.110.55 27,682 0.93 0.81 0.020 0.110.50 29,765 0.90 0.79 0.020 0.100.45 31,595 0.88 0.77 0.019 0.100.40 35,404 0.83 0.73 0.018 0.09CuEq

Cutoff ktons CuEq, % Cu, % Mo, % Ag, opt

1.00 15,907 1.33 1.20 0.026 0.070.75 41,561 1.03 0.95 0.017 0.050.60 79,995 0.86 0.79 0.014 0.040.55 97,517 0.81 0.75 0.013 0.030.50 115,218 0.76 0.70 0.012 0.030.45 131,157 0.73 0.68 0.011 0.030.40 151,686 0.69 0.63 0.011 0.03CuEq

Cutoff ktons CuEq, % Cu, % Mo, % Ag, opt

1.00 23,800 1.36 1.21 0.030 0.110.75 56,740 1.07 0.96 0.020 0.080.60 104,868 0.88 0.81 0.016 0.060.55 125,199 0.83 0.76 0.014 0.050.50 144,983 0.79 0.72 0.013 0.050.45 162,752 0.76 0.69 0.013 0.040.40 187,090 0.72 0.65 0.012 0.04

CuEqCutoff ktons CuEq, % Cu, % Mo, % Ag, opt

1.00 10,923 1.34 1.24 0.020 0.070.75 29,675 1.03 0.93 0.019 0.070.60 58,034 0.86 0.78 0.016 0.050.55 73,190 0.79 0.72 0.015 0.040.50 85,841 0.75 0.68 0.014 0.040.45 98,952 0.71 0.65 0.014 0.030.40 136,222 0.63 0.58 0.012 0.02

Inferred

Measured

Indicated

Measured + Indicated


SGS FORM No. A263a-2/14/13 !


14.8 Uncertainties

The resource estimates assume that the copper and molybdenum at Copper Creek occurs entirely in the form of sulfides. However, approximately 10 percent of the intervals at Copper Creek are assayed for acid soluble copper, and acid soluble/total copper ratio values in these intervals indicate that significant oxidation is present down to depths of at least 500 feet in the Globe, Copper Prince, Old Reliable and Peripheral areas, although oxidation in Childs-Aldwinkle, Mammoth and American Eagle is mostly superficial.

Table 14.11 summarizes mean acid soluble/total copper ratios by depth range and deposit. (Ratios of over 0.1 are indicative of some oxidation, ratios of over 0.2 are indicative of significant oxidation and ratios of over 0.5 are indicative of dominant oxidation. Blank cells signify no data.) This work is from the June 2012 technical report and has not been updated with the additions to the drill hole data base.

The current resources make no allowance for lower mill recoveries in oxidized or partially-oxidized material because the distribution of oxidation could not be defined in the model (the acid soluble copper data are not extensive enough and the assay data base contains no “oxide” variable). However, the Globe, Copper Prince and Old Reliable deposits contain only a small percentage of the total resource, so the impacts are unlikely to be appreciable, but should be kept in mind as the project moves forward.

Table 14.11 Acid Soluble to Total Copper Ratio, Assays, Total Copper Greater Than 0.10%

Area 0-100, ft 100-200, ft

200-400, ft

400-600, ft

600-800, ft

800-1000, ft

>1000, ft

1. Globe 0.69 0.26 0.28 0.28

2. Copper Prince 0.56 0.28 0.20 0.32 0.15

3. Old Reliable 0.42 0.37 0.25 0.18 0.22

4. Childs-Aldwinkle 0.33 0.16 0.16 0.06 0.08 0.04

5. Mammoth 0.21 0.08 0.08

6. American Eagle 0.24 0.08 0.11 0.04 0.04 0.02 0.07

0. Peripheral 0.53 0.65 0.25


SGS FORM No. A263a-2/14/13 !


15.0 MINERAL RESERVE ESTIMATES

No mineral reserve has been estimated.


SGS FORM No. A263a-2/14/13 !

! 117

16.0 MINING METHODS

16.1 Tonnage Estimate for Post Pillar Cut & Fill Mining

As a follow-up to a previous block caving study, a review of a more selective mining method for extracting the AE, Keel and Mammoth deposits was undertaken. The most effective mining method would allow extraction of the higher-grade portions of each deposit, at a somewhat lesser production rate, but still provide an acceptably low unit operating cost, deliver a higher head grade to the mill, and require less preproduction and sustaining capital. A review of the potential mining methods indicated that only a post-pillar cut and fill method would meet all of the requisite parameters.

Post-pillar cut and fill is a method similar to room and pillar mining, but allows the recovery of deposits that are more than 100 ft. thick. Mining proceeds from the bottom, upward. After each level is mined, it is backfilled with an engineered backfill. The pillars are designed to yield while mining the fourth or fifth lifts; so the pillar yielding occurs within the fill, thus limiting any movement.

MineSight mine planning software, bench maps and sections were used to identify higher-grade areas within the AE, Keel and Mammoth deposits, and the lower portion of the Childs Aldwinkle deposit, that could be mined by a post-pillar cut & fill method. As a basis for mine planning, four areas were selected in the Keel and American Eagle (AE), which could produce approximately 6,250 t/d each, for a total production rate of 25,000 t/d. Early production would include mining of the upper breccia deposits (Childs, Prince, and Globe) to take advantage of the tons and higher grades available from these deposits. In addition, it is assumed that the small Old Reliable open pit should also be included with this mine plan to provide additional tons and to provide a source for backfill that may be needed in the Keel/AE mining.

16.2 Cutoff Evaluation

Since this study would be classed as a “Prel iminary Economic Assessment” (PEA); “measured”, “indicated” and “inferred” resources were included in the evaluation. In order to calculate the portion of the various deposits that would be available for mining, it was necessary to identify the resources within those deposits that might be economically extracted.

The economic portion of the resource is typically determined by the application of a breakeven cutoff grade, or NSR value, that considers the total operating cost (mine, plant and administration), metal price(s), process recovery(s), applicable royalties, and forward costs for concentrate freight, insurance, smelting and/or refining.

These parameters are equated to determine the minimum grade, grade equivalent, or value of metal(s) that will need to be mined in order to cover the total direct operating costs.


SGS FORM No. A263a-2/14/13 !

! 118

The rock conditions in the AE and Keel deposits appear to be much better than parts of the breccia deposits, and the mineralization appears to be more consistent.

Although gold, silver and molybdenum are typically present in minor quantities in all of the deposits; only molybdenum and silver have been considered as payable bi-product metals in this evaluation. With three metals present, it was deemed easier to express the breakeven cutoff grade as a copper equivalent grade.

Since the breakeven cutoff grade represents the minimum grade, or value, that will be mined, the average grade, or value delivered to the mill, will always be higher than the breakeven cutoff grade. The increment, between the breakeven cutoff grade and the mill head grade, provides for return of the capital investment and profit on the investment.

Other cutoff grades (incremental) may be employed later in the mine planning process, by the mine planners/management, to handle situations where mineralized material, with a value below the economic cutoff grade, must be mined in order to reach ore, or to optimize the cash flow. However, these incremental cutoff grades are not normally used in determining the initial, breakeven cutoff grade used to establish minable material.

The following algorithm illustrates the typical relationship between the various parameters to calculate at the breakeven cutoff copper equivalent grade:

If the breakeven cutoff is expressed as a value, it is the sum of the mine, mill, G&A, and forward operating costs. If an NSR (net smelter return) is used (including concentrate freight and royalties), it must exceed the total on-site unit operating costs. The following parameters were used in determining the breakeven copper equivalent cutoff grade:

Production Rate (stpd): 25,000 Process Recovery: Copper (%) 0.890 Molybdenum (%) 0.840 Silver 0.500

Smelter Recovery (%) 0.965

Cueq. = Total Operating Cost

(Cu Price + Credits – Fwd. Costs) x Cu Rec. x 20 lbs/t/%


SGS FORM No. A263a-2/14/13 !

! 119

Metal Prices: Copper ($/lb) $2.75 Molybdenum ($/lb) $12.00 Silver $20.00

Unit Operating Costs ($/t): Mine $17.45 Mill 4.98 G&A 0.60 Subtotal Oper. Costs $26.05

Freight, Smelting & Refining: Truck/Rail Freight ($/t conc.) $12.00 Smelting ($/ton conc.) 81.65 Refining ($/lb Cu) 0.09

F, S & R ($/t milled material) $2.41

Royalty ($/lb Cu) $0.08

Bi-Product Credit: Molybdenum ($/lb Cu) $0.16 Silver ($/lb Cu) 0.07 Total Bi-Product Credit $0.23 Equating these parameters provides the following breakeven copper equivalent grade for a 25,000 stpd production rate.

The large, undiluted resource tonnage, and the relatively low copper grade, associated with a 0.55 Cueq. cutoff, indicated that the deposits would preferably be mined by a highly-productive, low operating cost mining method, such as block caving, or by a post-pillar cut and fill mining method. A preliminary study has indicated that mining the AE and Keel deposits using a block caving mining method, could not be economically justified. The use of a post-pillar cut and fill method allows for much more selectivity in the mining process, but limits the production from any one stoping area to less than 6,500 stpd, and results in a slightly higher mine unit operating cost.

The resource tons and grades used for the PEA mine plan contained within the AE, Keel, Mammoth and lower Childs deposit grade shells, at a 0.55 percent CuEq., were tabulated by Independent Mining Consultants (IMC), for each 20 foot bench height, between elevations 500 and 4200, with the results presented in the following

Cueq. =

US$ 26.05/t + US$2.41= 0.564% CuEq

(US$ 2.75 + US$ 0.23 – US$ 0.08) x 0.86 x 20#/t/%


SGS FORM No. A263a-2/14/13 !

! 120

Table 16.1. This preliminary economic assessment (PEA) is preliminary in nature and includes inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves. There is no certainty that the preliminary economic assessment will be realized. Mineral resources that are not mineral reserves do not have demonstrated economic viability.

A new 20 ft x 20 ft x 20 ft model was developed by IMC based on the latest drilling input, and new tonnages and grades tabulated. Bench plans on 20 ft centers were drawn to indicate above cutoff areas in the Keel and AE deposits.

Using MineSight software, four areas, above the 0.55 percent Cueq. cutoff grade, were then identified, in the AE, Keel, Mammoth and lower Childs deposits, that contained a sufficient tonnage, which could be mined by a post-pillar cut and fill mining method, and each could support a production rate of 6,250 t/d. Haulage level elevations were then located at the base of each area.

The individual bench maps within the deposit grade shells were then reviewed for above cutoff grade blocks that would not support development access, or needed to be re-classified by location. These changes were then tabulated to reflect the minable tons and grades available within the resource shells. Factors for mining (external) dilution and recovery, based on the proposed mining method, were applied. The results of these exercises indicate the potentially minable tons and grades within the resource grade shells for the various deposits, and were used in the cash flow evaluation of the Copper

Table 16.1 Post-Pillar Cut & Fill Mineralized Material

Classification Deposit Tons Grade (%Cueq.)

Contained Klbs Cu

Indicated American Eagle 84,364,000 0.76 1,282,352

Indicated Keel 40,665,000 0.75 609,975

Measured/Indicated Mammoth 4,547,000 1.32 120,041

Indicated Childs Prince/Globe 2,464,000 1.66 81,805

Total 132,040,000 0.79 2,094,173

Inferred American Eagle 55,024,000 0.74 814,355

Inferred Keel 17,276,000 0.66 228,043

Inferred Mammoth 527,000 1.03 10,856

Inferred Childs Prince/Globe 1,244,000 1.60 39,808

Total 74,071,000 0.74 1,093,062


SGS FORM No. A263a-2/14/13 !

! 121

Creek deposits. The following Table 16.2 presents a summary of the results which makes up the PEA mine production schedule. The mine plan of this (PEA) is preliminary in nature and includes inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves. There is no certainty that the preliminary economic assessment will be realized. Mineral resources that are not mineral reserves do not have demonstrated economic viability.

16.3 Geotechnical Considerations

Call & Nicholas, Inc. (CNI) has evaluated the American Eagle, Keel, Mammoth and Childs deposits as potential post-pillar cut and fill operations (PPCF). The American Eagle and Keel deposits have similar RQD distributions and intact strength properties which are greater than those of the Mammoth and Childs.

Based on surface geologic structures, a room width of 20 ft is recommended using 4-ft long split set bolts on 3 to 4 ft centers. As additional joint set data from within the deposits are gathered, there is the potential for the American Eagle and Keel to be

Table 16.2 Diluted/Recovered Post-Pillar Cut & Fill Mineralized Material*

Classification & Deposit KTonsGr.

(%Cu)Gr.

(%Mo) Ag (oz/t) KLbs Cu

Indicated

American Eagle 63,905 0.76 0.012 0.03 971,356

Keel 30,804 0.75 0.025 0.12 462,060

Mammoth 3,026 1.27 0.003 0.12 76,860

Chiles/Prince/Globe 2,013 1.65 0.046 0.13 66,832

Total 99,748 0.790 0.016 0.063 1,577,108

Inferred

American Eagle 41,681 0.74 0.012 0.03 603,031

Keel 13,086 0.66 0.018 0.10 167,179

Mammoth 338 1.03 0.004 0.10 7,202

Childs/Prince/Globe 1,037 1.40 0.021 0.08 28,969

Total 56,142 0.735 0.014 0.048 806,381

*Extensions may vary due to rounding


SGS FORM No. A263a-2/14/13 !

! 122

widened to a 25 ft room width, while in the Mammoth and Childs room widths may need to be reduced.

The recommended pillar widths for the American Eagle and Keel are 20 by 20 ft, and in the Mammoth and Childs 25 by 25 ft. These pillar widths are based on mining 15ft high lifts for five (5) lifts after which mining heights can be increased to 20 ft.

Because the Mammoth and Childs deposits are at, or near, surface and because the actual mining area is yet to be finalized, CNI used a 70 degree crack angle to identify where Copper Creek is likely to be impacted and where to locate potential shaft locations.


SGS FORM No. A263a-2/14/13 !

! 123

CNI recommends diverting Copper Creek when either of the following happens:

a) The fill height of the American Eagle, or Keel, is 600 ft, or b) When post pillar cut and fill mining starts at either the Mammoth or Childs.

The mine schedule includes open stope mining part of the Childs and Mammoth deposits as the access is developed down to the American Eagle and Keel. This mining will be of a limited area and the backfill will be a cemented fill. Therefore, the surface subsidence is unlikely.

The location of potential shaft sites based on the property boundary and estimated crack limit can be seen in Figure 16.1. The predicted distance over which Copper Creek should be diverted is shown in Figure 16.1.

Shaft location Option 1 is preferred because it is at a lower elevation and further away from the proposed Old Reliable pit crest.

CNI's full report and details supporting these geotechnical recommendations can be found in Appendix 2.


SGS FORM No. A263a-2/14/13 !


! Figure 16.1 - Deposits with 70 Degree Crack Limits


SGS FORM No. A263a-2/14/13 !


16.4 Basic Mine Plan

The basic mine would access the Mammoth-Keel and AE deposits mine through a 33 ft diameter, 3,800 ft deep, vertical shaft located at coordinates 638765 N and 401629 E. These coordinates may change with better definition of the deposits, surface topography, and Redhawk “patented” property boundaries. The current location was the result of evaluating four different locations, which considered the surface road access, deposit locations, Redhawk “patented” property boundaries, Old Reliable pit limits, and the projected geotechnical crack lines to the surface. This shaft size is capable of handling a production rate of 25,000 tons per day.

Four major production levels (16 ft x 16 ft in Section) were selected. These levels would allow access from the shaft to the base of the most favorably-mineralized areas in the Mammoth, Lower Childs, Keel and AE deposits. The uppermost level (3000 Level) would access the bottom on the Mammoth and Lower Childs deposits, while the 1000 and 1500 Levels would access both the Keel and AE deposits. The lowest level (500 Level) would access the deeper extensions of the Keel deposit only.

All four haulage level shaft stations are planned to be developed during the shaft sinking phase to allow level development on the 500, 1000, 1500, and 3000 levels.

Internal ramps (15 ft x 15 ft), from the major haulage levels, would allow access to successive mining lifts and the base of the secondary deposits located above the haulage levels. The initial production would be situated in four stoping areas – two in the Keel (one above the 1000 Level, and one above the 1500 Level), and two in the AE deposit – (one above the 1000 Level, and one above the 1500 Level). Each stope would be designed for a production rate of 6,250 tons per day, for a total production rate of 25,000 tons per day.

Two, 20 ft diameter, approximately 3,000 ft deep, ventilation shafts would be sunk from the surface to the lowest level in each of the Mammoth-Keel and AE deposits, with connections to each of the haulage levels. Ten foot diameter, bored, rock and ventilation passes would connect the mining levels with the haulage levels.

Mining would proceed from the haulage level upwards in 20 ft lifts; utilizing a post-pillar cut and fill mining method (see Figure 16.2). Blast-hole drilling would utilize electric-hydraulic, two-boom, jumbos. Rock loading would be performed with 9 to10 cy LHD’s, which would move the blasted rock from the faces, or muck bays, to rock passes that would transfer the blasted rock to the haulage level. Fifty-ton haulage trucks would move the rock from the passes to grizzlies, equipped with a rock breaker, located on each haulage level, approximately 1,300 ft from the shaft. The broken rock would pass through the grizzlies to 60-inch jaw crushers, be reduced to a nominal 6-inch product, and then be moved via an inclined conveyor to a 4,000 ton storage pocket located adjacent to the production shaft. Fifty-five ton skips would move the rock through the shaft to the surface (Figure 16.3).


SGS FORM No. A263a-2/14/13 !


When a 20 ft lift has been exploited, bulkheads will be installed to contain an engineered backfill consisting of de-slimed mill tailings, classified mine waste, fly ash, and cement. When this backfill has consolidated, the next lift can be started on the top of the previously-excavated lift. The use of an engineered backfill will allow smaller pillars than would normally be allowed, to be left. This will optimize mining recovery of material from each of the deposits. In addition, by returning a large portion of the mill tailings back into the mine, the surface tailings impoundment size, and cost, can be minimized.

In general, the rocks at Copper Creek are very strong, with an average compressive strength of 10,000 to 20,000 psi, and a median RQD’s in the range of 80 to 90 percent. The phreatic surface has been interpreted to be at the level of the San Pedro river, which is around the 2,000 ft elevation. The surface elevation over the deposits is around the 4,200 ft elevation, with the upper limit of reasonably contiguous mineralization in both deposits being around the 2,000 ft elevation. The better mineralization, in the both deposits, continues downward to approximately the 500 ft elevation. Preliminary geotechnical investigations indicate that rooms 20 ft wide and 20 ft high can be excavated in the rocks in the Mammoth-Keel and AE deposits, leaving 20 ft square pillars for initial support. This would result in a mining recovery of 75 percent. The RQD’s in the breccias (Mammoth/Childs) are somewhat less, thusly CNI has recommended leaving 25 ft x 25 ft pillars in these areas, resulting in an overall recovery there of 65 percent.

The upper (above the 3500L) Childs, Prince, and Globe breccias deposits would be developed and mined by the previously-proposed decline and access drifts, with a cut and fill mining method. A third, 16 ft diameter, ventilation shaft, as previously discussed, would be located adjacent to the upper Childs deposit access.


SGS FORM No. A263a-2/14/13 !


! Figure 16.2 – Mining Method Schematic


SGS FORM No. A263a-2/14/13 !


! Figure 16.3 Typical Level/Shaft Section


SGS FORM No. A263a-2/14/13 !


16.5 Preproduction Infrastructure/Development/Equipment

The underground preproduction phase includes all surface and underground preparation necessary to initiate production. Typically, this includes: site preparation, mine surface infrastructure construction, utilities, production and ventilation shaft sinking, emergency escape-way system installation, equipment purchases, and the underground utilities, infrastructure and development necessary to initiate production.

During the preproduction period, the Company infrastructure will be limited to administration/quality control office trailers, while the permanent surface infrastructure is being constructed.

All preproduction construction and underground development is assumed to be performed by a mining contractor. During the preproduction phase, a training program will be initiated that will allow company employees to perform all production-related tasks once the preproduction phase is completed.

The following Table 16.3 summarizes the estimated preproduction quantities and construction times.

Table 16.3 Preproduction Development Quantities & Construction Time Estimates

Cost Center Unit Quantity Construct. Time, mo.

Shaft Site Preparation Lump Sum 1 3

Production Shaft (33 ft dia.) V.F. 3,800 24

Shaft Stations (4) CuFt each. 42,400 3

Shaft Change-over Lump Sum 1 3

1000L Level Devel. (16 ft x 16 ft) L.F. 7,000 16

1500L Level Devel. (16 ft x 16 ft) L.F. 7,000 16

Vent Shaft #1 (20 ft dia.) V.F. 3,000 18

Vent Shaft # 2 (20 ft dia.) V.F. 3,000 18

1000L Vent. Drifting (16 ft x 16 ft) L.F. 300 1

1500L Vent. Drifting (16 ft x 16 ft) L.F. 500 1

1000L Access Drifting (15 ft x 15 ft) L.F 500 2

1500L Access Drifting (15 ft x 15 ft) L.F 500 2

1000L Maint. Area (16 ft x 16 ft) L.F. 1,970 8

1500L Maint. Area (16 ft x 16 ft) L.F. 1,970 8


SGS FORM No. A263a-2/14/13 !


During the preproduction period, the Company equipment will be limited to pickup/SUV trucks and office/engineering/monitoring equipment. Company personnel during the preproduction phase includes: project management, quality and environmental control, mine planning engineers and Company personnel and purchasing managers. The following Table 16.4 lists the anticipated Company personnel during the preproduction phase.

The following Figure 16.4 presents the estimated preproduction development schedule.

1000L Crusher Exc./Install CuFt 180,000 10

1500L Crusher Exc./Install CuFt 180,000 10

1000L Conveyor Incline (15 ft x 15 ft) L.F 1,200 6

1500L Conveyor Incline (15 ft x 15 ft) L.F 1,200 6

1000L Station Pocket Exc. CuFt 134,000 4

1500L Station Pocket Exc. CuFt 134,000 4

Table 16.4 Preproduction Personnel

Classification Number

Project Manager 1

Environmental Manager 1

Personnel Manager 1

Purchasing Manager 1

Engineers 3

Geologists 1

Secretary 2

Clerks 1

Janitors 1

Total Personnel 12


SGS FORM No. A263a-2/14/13 !


Figure 16.4 – 25,000 tpd Preproduction Development Schedule

Month→ 1 2 3 4 5 6 7 8 910 11

12

13

14

15

16

17

18

19

20

21

22

23

24

25

26

27

28

29

30

31

32

33

34

35

36

37

38

39

40

41

42

43

44

45

464748

Cost CenterBreccia Decline to 3000L Drift to Prince & Globe Deposits Acc./Prep. Childs, Prince, Globe

Shaft Site Preparation Mine Surface Infrastructure Prod. Setup/Shaft Sinking

Shaft Stations (4)

Shaft Changeover 1000L Level Development 1000L Vent Access Drifting

Vent Shaft #1

1000L Maint Area 1000L Crusher Exc./Install 1000L Conveyor Incline

1000L Storage Pocket

Vent Shaft #2 1500L Development (as 1000L) 3000L Development (as 1000L) 500L Development (as 1000L) Miscellaneous Exc./Install’s


SGS FORM No. A263a-2/14/13 !

! 132

16.6 Production

16.6.1 Production Personnel

Direct and indirect labor requirements were established to match the selected mining method, support systems, equipment and general mine requirements for a shaft-accessed, 25,000 stpd, and underground mining operation. Personnel requirements are based on an operating schedule of 8 hours per shift, three shifts per day, seven days per week. The mine is scheduled to operate 360 days per year. Many new mines are now working on a two, 12 hour shift basis, which may prove to be a more effective schedule once more detailed engineering is completed.

It is understood that training of the work force will be an important requirement for the success of the project. However, this study assumes a trained workforce will be in place.

Personnel are presented as direct and indirect staffing for the production phase of the project. All preproduction development work is assumed to be performed by a reputable underground mining contractor.

The Owner’s project and quality control team will oversee the work performed by the Contractor, and coordinated by the EPCM Contractor, during the preproduction period.

The following Table 16.5 lists the estimated direct and indirect mine labor for staffing a 25,000 stpd post-pillar cut and fill underground mining operation.


SGS FORM No. A263a-2/14/13 !

! 133

Table16.5 Personnel Summary

Classification Working/Day Rotated Off/Day

Total on Payroll

Direct

Miners 60 20 80

Miners Helpers 60 20 80

Equipment Operators 110 36 146

Backfill Operators 24 8 32

Hoist-men 3 1 4

Top-landers 3 1 4

Mechanics 24 8 32

Mechanics Helpers 24 8 32

Electricians 5 1 6

Electrician Helpers 4 1 5

Definition drillers 6 2 8

Materials Handling 27 9 36

Labor 18 6 24

Subtotal Direct 368 121 489

Indirect

Mine Manager 1 0 1

General Foremen 3 1 4

Shift Bosses 9 3 12

Chief Engineer 1 0 1

Engineers/Planners 3 0 3

Surveyors 2 0 2

Chief Geologist 1 0 1

Geologists 3 1 4

Samplers 6 2 8

Mine Office 3 0 3

Mine Dry 3 1 4

Subtotal Indirect 35 9 46

TOTAL 403 130 533


SGS FORM No. A263a-2/14/13 !

! 134

With the mine scheduled at three shifts per day, seven days per week, and 24 hours per day, a fourth crew will be required to relieve rotating labor. This additional, non-working labor will require additional mine office and dry facilities, and add to the cost of accounting for this additional labor. The work schedule for this rotation schedule adds approximately 6.35 percent in scheduled overtime to the mine labor costs.

16.6.2 Shafts

The main production shaft will provide for essentially all of the movement of rock, personnel, materials, equipment, and supplies for the mine. It will also be the primary fresh-air ventilation intake. The shaft is a 33-ft diameter, concrete-lined shaft, equipped with steel sets on 20 foot centers. A 9 ft x 12 ft, counter-weighted, single deck cage will handle personnel, equipment, and supplies in and out of the mine. A small (10-man capacity) “auxiliary-chippy” cage is included to handle small groups of personnel entering and leaving the mine.

Because of the requirement for multi-level hoisting, all hoists are assumed to be ground-mounted units for moving skips, the main service cage and the auxiliary man (“Chippy”) cage. The production hoist will raise 55-ton skips, moving on rope guides, at speeds of approximately 3,000 fpm. The hoisting capacity from a depth of 3,800 feet is approximately 27,000 tpd.

The main service hoist will have a 40 ton payload capacity, travel at 2,000 fpm, and accommodate 150 personnel.

From a 4,000 ton bin located adjacent to the shaft, crushed material will be fed by apron feeders onto weigh belts that load the shaft skips on each haulage level. The belts will load the skips directly, with a separate belt for each skip.

On the surface, the skips will discharge into a 2,000 ton steel-lined bin. Material from the bin will be discharged by an apron feeder to an underlying conveyor that will carry the material to a stacking conveyor. A reclaim system from the stacker stockpile will feed an overland conveyor for movement of the material to the coarse storage located adjacent to the mill.

16.6.3 Maintenance Shops

Primary maintenance shops will be located on each haulage level adjacent to the # 1 exhaust ventilation shaft to allow for rapid exhausting of smoke and fumes in case of a fire in this area. The number of service bays is estimated from the fleet size and the equipment availability. The shops will be equipped to service diesel LHD’s, jumbos, trucks, rock bolters, and support equipment. A wash bay will be positioned so that all equipment can be washed prior to service. A 12-ton crane will be located in the access


SGS FORM No. A263a-2/14/13 !

! 135

drift, and smaller cranes located in each service bay. A dedicated sump will be used to collect mud carried into the shop area by the vehicles.

A dispatch office, lunch room, first aid and toilets will be located within the shop complex. Field service and repair trucks will be available for minor repairs and equipment lubrication.

A fuel storage station and storage tanks will be located in the shop area and fed by a borehole from the surface. Fuel will be dispatched daily from the surface in batches equal to the daily usage.


SGS FORM No. A263a-2/14/13 !

! 136

16.6.4 Explosives Storage

The underground storage facility will be constructed to accommodate ANFO, caps and high explosives for boosters, primers and stick powder. The explosives will be transported from the surface magazines via the shaft to the main underground magazine. Upon arriving at the underground shaft station, specially equipped and authorized transport trucks will deliver the explosives to the magazine by certified personnel. These trucks will be equipped with special warning signs and flashing lights on the front and rear. All other traffic in the travel way will be stopped, while the explosives are being transported.

The explosives will be stored in separate bays, separated by 60 feet of rock. The access to each storage bay will be secured with 7-foot high, locked gates, with two entrances provided for each bay. The location of the explosive facility will be a minimum of 300 feet from any blasting area, and at least 75 feet from the main travel way.

16.6.5 Compressor Requirements

The preliminary design will not include a main surface compressor station feeding compressed air through pipes in the supply shaft to the production areas of the mine. Rather, a combination of portable and equipment on-board compressors will provide the required compressed air for local requirements. Small, fixed compressors will be located in the underground maintenance shops.

16.6.6 Service Water Requirements

Mine service water will be supplied from the surface, as the mine is expected to be essentially dry. Mine service water will be needed for production and development drilling, wash-downs, and dust suppression. Peak usage rates are estimated to be approximately 675 gpm. The following Table 16.6 summarizes the estimated water requirements.


SGS FORM No. A263a-2/14/13 !

! 137

For a 19.5 hour day, this would equate to approximately 790,000 gallons per day. Waste water collection and clarification could reduce this amount by potentially 25 percent, or by approximately 200,000 gallons per day.

This total average water gpm requirement, over a 21 hour period, would equate to a total daily use of approximately 800,000 gallons per day in the underground mine. About 80 percent of the total gpm required, or approximately 500 gpm, would need to be clean water brought into the mine from the surface for drills, gland water, etc. The remaining 150-200 gpm, primarily needed for dust suppression, could come from water clarified within the mine.

Water collected from face drilling and ditches should be clarified in underground sumps and potentially recover about 20 percent of the 500 gpm clean water entering the mine, or about 100 gpm. In addition, the backfill coming into the mine from the mill will contain about 30 percent water. It has been assumed that one-half of this water would be decanted and clarified in the underground sumps and be available for dust suppression on roads, muck piles, crushers, etc. The decanted and clarified water from the backfill would amount to about 250 gpm, for a total of approximately 350 gpm available for dust suppression. With a need for only 150 to 200 gpm for dust suppression, this would indicate that only about 200 to 300 gpm of clarified water would need to be pumped from the mine.

Table 16.6 Estimated Service Water Usage

Description Fleet Qty. gpm/unit gpm consumed Utilization, % Est. Ave. gpm

Jumbos 18 30 540 80 432

Bolters 2 15 30 60 18

Misc. Dust Sprays 20 5 100 40 40

Diamond Drilling 2 8 16 75 12

Stopers/Jacklegs 10 3 30 25 8

Wash Bays 2 10 20 20 4

Conveyor Dust 4 3 12 80 10

Crusher Cooling 2 2 4 80 4

Gland Water 4 5 20 100 20

Roads 2 20 40 50 20

Miscellaneous 1 20 20 100 20

Contingency (15%) 1 88 88 100 88

Total 676


SGS FORM No. A263a-2/14/13 !

! 138

16.6.7 Power and Communications Systems

A primary voltage of 13.8 kV will be routed from the surface substation through the production shaft to underground primary substations located near the shaft station on each level. From there the voltage will be transformed to 4.16 kV for distribution throughout the level. Underground secondary transformer stations will provide feeds for crusher and conveyor requirements. The development and production loads will be fed by mine load centers that will provide 440V/220V/110V power to mobile equipment and temporary loads.

The total estimated power consumption by the mine is summarized in the following Table 16.7.

Communication throughout the mine will be provided by leaky feeder radio and VOiP telephone secondary system. A personnel emergency dispatch system (PEDS) will allow one-way, mine-wide, emergency communication from the surface to all cap lamps equipped with the PEDS pager.

Table 16.7 U.G. Power Requirements

Description No. kWh/UnitUtilization,

% Total kWh

Production Hoist 1 12,000 80 9,600

Service Hoist 1 5,000 40 2,000

Chippy Hoist 1 500 20 100

Jumbos 16 150 85 2,040

Raise Borer 1 215 20 43

Rock Breaker 2 95 50 95

U.G Crusher 2 160 90 288

U.G. Compressors 4 375 60 900

U.G. Conveyors 2 200 90 360

Feeders 4 125 90 450

Primary Pumps 2 200 100 400

Secondary Pumps 24 15 100 360

Primary Fans 2 200 100 400

Secondary Fans 24 15 100 360

Safety Equip./Lamps 1 20 100 20

Office Equipment 1 20 100 20


SGS FORM No. A263a-2/14/13 !

! 139

16.6.8 Ventilation

The total estimated quantity of ventilation air is estimated at 3,036,000 cfm, based on the equipment and personnel working within the mine, and summarized in Table 16.8. The basic ventilation scheme is to draw air down the production shaft and access decline to the upper Childs workings, pass the fresh air through the active mining areas and exhaust the contaminated air through three, concrete-lined, ventilation shafts fitted with primary exhaust fans on the surface. Within the mine active areas, ventilating air will be regulated with small axivane fans, stoppings and air doors.

U.G. Communications 1 20 100 20

U.G. Monitoring System 1 10 100 10

Total 17,066


SGS FORM No. A263a-2/14/13 !

! 140

16.6.9 Production Schedule

Table 16.8 Ventilation Requirements

Description No. Hp/Unit Total Hp Utilization, % Design Hp

Equipment

Jumbos 16 160 2,560 75 1,920

7 cy LHD’s 20 324 6,480 80 5,184

3 cy LHD’s 4 293 1,172 60 703

Haulage Trucks 16 572 9,152 80 7,321

Bolters 2 160 320 60 192

Scalers 2 160 320 60 192

U.G. Grader 2 320 640 50 320

Backfill Trucks 3 206 618 80 494

Explosives Trucks 2 147 294 75 221

Lube Trucks 2 160 320 60 192

Shotcrete Units 2 160 320 50 160

Fire Trucks 2 160 320 10 32

Water Trucks 2 160 320 50 160

Scissors Lift Trucks 2 128 156 60 94

Personnel Carriers 2 147 294 30 88

Boss Buggies 8 85 680 80 544

Total Equipment 23,966 74.3 17,817

CFM @ 125 cfm/Hp 2,225,000

Personnel

150 ea. @ 500 cfm/ea. 75,000

Subtotal, CFM 2,300,000

Recirculation @ 20%, CFM 460,000

Fan Efficiency @ 90%, CFM 276,000

Total Ventilation, CFM 3,036,000


SGS FORM No. A263a-2/14/13 !

! 141

The full production rate of 25,000 t/d is scheduled to be produced primarily from four separate working areas within the Keel, and American Eagle deposits, with some contribution from the breccias in the early years to enhance the mill feed grade. Each area will normally be programmed to produce 6,250 t/d, with variations as dictated by mine planning. The mine is planned to operate three shifts per day, seven days per week, 360 days per year. Initial production is scheduled from the 1000 and 1500 levels in the Keel and American Eagle deposits, with two stoping areas in each deposit. The following Table 16.9 provides a summary of the projected production tons and grades for the anticipated 18-year mine life. The first year production reflects a 20 percent reduction from the scheduled full production rate to allow for a start-up learning curve.


SGS FORM No. A263a-2/14/13 !

! 142

The mine plan of this (PEA) is preliminary in nature and includes inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves. There is no certainty that the preliminary economic assessment will be realized. Mineral resources that are not mineral reserves do not have demonstrated economic viability.

Table 16.9 REDHAWK RESOURCES

Year KTons Cu, % Mo, % Oz Ag/t

1 7,200 0.95 0.018 0.09

2 9,000 0.89 0.019 0.10

3 9,000 0.92 0.019 0.08

4 9,000 0.74 0.015 0.06

5 9,000 0.73 0.014 0.06

6 9,000 0.73 0.014 0.06

7 9,000 0.73 0.014 0.07

8 9,000 0.73 0.015 0.07

9 9,000 0.73 0.015 0.07

10 9,000 0.71 0.023 0.07

11 9,000 0.75 0.012 0.03

12 9,000 0.75 0.012 0.03

13 9,000 0.75 0.012 0.03

14 9,000 0.75 0.012 0.03

15 9,000 0.73 0.021 0.03

16 9,000 0.77 0.017 0.05

17 9,000 0.79 0.011 0.03

18 4,690 0.73 0.014 0.025

Total 155,890 0.77 0.015 0.05


SGS FORM No. A263a-2/14/13 !

! 143

17.0 RECOVERY METHODS

17.1 Process Description

The design criteria developed for this study is attached as Appendix 4. The operation is designed to process 25,000 dry tons per day. The design basis for the various unit operations is included in the Design Criteria.

Primary crushing, SAG and ball mills, bulk rougher and cleaning flotation, differential coppermolybdenum flotation and concentrate filtering and handling operations are planned to process the design tonnage.

The process plant is designed to operate 24 hours per day, 7 days per week and 360 days per year. The utilization factor used for the calculation of the nominal hourly flow rates is 92.5 percent. Metallurgical work indicates that copper recovery increases at finer grind sizes. For the purposes of this study a grind size of 80 percent passing (P80) of 140 microns has been specified.

Process flowsheets developed for the scoping study are attached in Appendix 5. Flowsheets developed are summarized below:

The equipment list developed for the project is attached in Appendix 6.

10-F-01 Flowsheet, Coarse Ore Storage

15-F-02 Flowsheet, Grinding

40-F-06 Flowsheet, Rougher Flotation

40-F-07 Flowsheet, Regrind

40-F-08 Flowsheet, Copper – Moly Cleaner Flotation

45-F-09 Flowsheet, Moly Flotation

45-F-10 Flowsheet, Moly Flotation

50-F-11 Flowsheet, Copper Concentrate Handling

55-F-13 Flowsheet, Moly Concentrate Handling

60-F-30 Flowsheet, Reagents

60-F-31 Flowsheet, Reagents

70-F-35 Flowsheet, Tailings Handling

80-F-40 Flowsheet, Process & Reclaim Water

90-F-45 Flowsheet, Raw Water


SGS FORM No. A263a-2/14/13 !

! 144

17.2 ROM Handling

The crushing circuit will consist of single-stage primary jaw crushers that will be located underground in the mine. The crushers reduce the size of run-of-mine material from maximum 18 inches to nominally 80 percent passing 5 inches. Crushed rock is transported to the surface from underground and drops into a dump hopper. An apron feeder transfers crushed rock at a controlled rate from the dump hopper to an overland conveyor. Dust is controlled in the dump pocket and transfer points with water sprays.

Primary crushed material will be stockpiled on the ground. One reclaim tunnel will be installed beneath the stockpile. The stockpile will contain approximately 25,000 tons of live storage. When required, material will be moved from the dead storage area to the live storage area by front-end loaders or bulldozers.

Material will be withdrawn from the coarse reclaim stockpile by variable speed apron feeders (two operating and one standby). The feeders will discharge to a conveyor belt. Conveyor will discharge to the SAG mill in the grinding circuit. The reclaim rate will be monitored by a belt scale mounted on the conveyor. Crushed pebbles will be added at the SAG Mill Feed Conveyor.

Dust control in the Coarse Ore Stockpile Area will be by the dry type dust collector system. It will be installed to control dust in the reclaim tunnel.

17.3 Grinding Plant

Material will be ground to rougher flotation feed size in a semi-autogenous (SAG) mill primary grinding circuit and a ball mill secondary grinding circuit.

The SAG mill will operate in closed circuit with a trommel screen, pebble wash screen, and a pebble crusher. Trommel screen undersize will flow by gravity to the cyclone feed pump box. Trommel screen oversize will be transported by conveyors to the pebble crushing system. Tramp iron and broken media will be removed using two self-cleaning belt magnets that will be installed over the pebble conveyor ahead of the pebble crusher. A pebble wash screen, a pebble wash pump box and pebble wash pumps will be installed between the two pebble conveyors to wash the pebbles. The pebble wash pumps will discharge the pebble wash screen undersize to the cyclone feed pump box.

The pebble crushing circuit will consist of a 200 ton SAG oversize surge bin, pebble crusher belt feeder and a short-head cone type crusher. Trommel screen oversize can be bypassed around the pebble crushing circuit via diverter gate ahead of the pebble crusher to the pebble conveyor which feeds the SAG mill feed conveyor. Trommel screen oversize will normally discharge from the transfer conveyor through the diverter gate into a surge bin ahead of the pebble crusher. Material will be removed from


SGS FORM No. A263a-2/14/13 !

! 145

the surge bin by variable speed belt feeder that will feed the pebble crusher. The crusher will discharge onto the pebble conveyor. A belt scale mounted on a conveyor will monitor the SAG mill recycle rate, and another belt scale mounted on a conveyor will monitor the pebble crusher feed rate to the SAG mill conveyor.

Secondary grinding will be performed in two ball mills operated in parallel. Each ball mill will operate in closed circuit with hydrocyclones. Ball mill discharge will be combined with trommel screen undersize in a cyclone feed pump box and will be pumped to hydrocyclone clusters. Combined slurry will be pumped using variable speed horizontal centrifugal slurry pumps (two operating and one uninstalled spare) to the primary cyclone clusters. Hydrocyclone underflow will flow by gravity to the ball mills. Hydrocyclone overflow (final grinding circuit product) will flow by gravity to the flotation circuit.

Cyclone overflow from each cyclone cluster will be sampled by primary samplers and analyzed by the copper on-stream analyzer for metallurgical control prior to flotation. Cyclone overflow from each cyclone cluster will also be monitored for particle size distribution by particle size monitor.

Grinding balls will be added to the SAG mill by a ball loading system. Grinding balls will be added to ball mills by a ball mill ball loading system.

Air compressors and instrument air dryer will provide service and instrument air for operation and maintenance.

An overhead crane will be installed for maintenance of the grinding mills.

17.4 Flotation Plant – Bulk (Copper-Moly) Flotation

Hydrocyclone overflow will flow by gravity to the bulk (copper-moly) flotation circuit. The copper-moly flotation circuit will consist of rougher/rougher scavenger flotation cells, concentrate regrind circuits, copper first cleaner/first cleaner scavenger flotation cells and re-cleaner flotation cells. The rougher row will consist of ten tank-type rougher flotation cells (four rougher flotation cells and six rougher scavenger cells). The rougher flotation concentrate from the first four cells in the rougher flotation row will be sampled with the rougher concentrate primary sampler and pumped by the rougher concentrate pump (one operating and one stand-by) to the regrind cyclone feed pump-box. The flotation concentrate from the last six cells in rougher flotation row will be pumped back to rougher flotation feed tank. Tailings from rougher flotation row will be sampled. The rougher flotation tails will flow by gravity to the tailing de-aeration tank. Copper-moly concentrate regrinding will be performed in a vertical mill. The vertical mill will operate in closed circuit with hydrocyclones. Vertical mill discharge will be combined with copper-moly rougher flotation concentrate and cleaner scavenger flotation concentrate in the copper regrind cyclone feed pump box and will be pumped by a variable speed horizontal centrifugal slurry pump (one operating and one stand-by) to a


SGS FORM No. A263a-2/14/13 !

! 146

hydrocyclone cluster. Hydrocyclone underflow will feed the regrind mill which discharges back to the copper regrind cyclone feed pump box. Hydrocyclone overflow (final regrind circuit product) will flow by gravity to the copper first cleaner flotation circuit.


SGS FORM No. A263a-2/14/13 !

! 147

Reground rougher and cleaner scavenger concentrate will feed four (4) copper first cleaner flotation cells. Concentrate from the copper first cleaner flotation cells will flow by gravity to the copper cleaner concentrate pump box. Tailings from the first cleaner cells will feed six (6) copper first cleaner scavenger flotation cells. Concentrate from the copper first cleaner scavenger flotation cells will flow by gravity to the copper cleaner scavenger concentrate pump box. Tailings from the cleaner scavenger cells, sampled by the scavenger tail primary sampler for process control, will flow by gravity to the tailings de-aeration tank.

First cleaner flotation concentrate will be pumped (one operating and one stand-by) from the cleaner concentrate pump box to the cleaner/re-cleaner feed distributor to feed either copper re-cleaner flotation cells or cleaner flotation cells. Concentrate from the re-cleaner flotation cells flow by gravity to the re-cleaner concentrate pump box, will be sampled with samplers and pumped by a re-cleaner concentrate pump (one operating and one stand-by) to the bulk concentrate thickener. Tailings from the re-cleaner flotation cells will feed to copper first cleaner flotation cells.

Air compressors, air receivers, and instrument air dryer will be installed for general plant operation and maintenance.

A bridge crane will be installed for maintenance of the flotation and regrind equipment.

Flotation reagents will be added at several points in the copper-moly flotation circuit.

17.5 Molybdenite Flotation

Copper-moly concentrate will report to the bulk concentrate thickener. The copper-moly thickener overflow will flow by gravity to thickener overflow manifold which will discharge to the process water pond. The bulk concentrate thickener underflow will be pumped by a variable speed horizontal centrifugal slurry pump (one operating and one stand-by) to a surge tank at the moly plant, the surge tank will serve to buffer surges for the moly plant. Sodium hydrosulfide (NaHS) will be used as the depressant for the copper minerals.

The molybdenite flotation circuit will consist of one row of molybdenite separation (rougher) flotation cells, one row molybdenite first cleaner flotation cells, and one molybdenite second, third, fourth, and fifth cleaner column cells.

The molybdenite separation (rougher) flotation row will consist of six molybdenite separation (rougher) flotation cells. Concentrate from the molybdenite separation cells will be delivered to the molybdenite rougher concentrate pumpbox and then pumped to moly first cleaner flotation circuit. Tailings from the molybdenite rougher cells will be pumped to the copper concentrate thickener.


SGS FORM No. A263a-2/14/13 !

! 148

The moly first cleaner flotation circuit will consist of five (5) flotation cells. Concentrate from the molybdenite first cleaner cells will be pumped by a froth pump (one operating/one spare) to the moly second cleaner column cell. Tailings from the molybdenite first cleaner flotation cells will be pumped to the moly rougher flotation circuit or to the copper concentrate thickener. In order to reduce consumption of NaHS, the rougher and first cleaner cells will be hooded and the flotation gas will be recycled. The moly second cleaner flotation circuit will consist of a column flotation cell whose concentrate will flow by gravity to the moly third cleaner column flotation cell, and tailings will be pumped by an air operated diaphragm pump (one operating/one spare) to the moly middling thickener.

Moly third cleaner concentrate will flow by gravity to the moly fourth cleaner column cell, and tailings will be pumped by an air operated diaphragm pump (one operating/one spare) to the moly second cleaner column cell.

Concentrate from the moly fourth cleaner flotation column cell will flow by gravity to the moly fifth cleaner flotation column cell. Tailings from the moly fourth cleaner flotation column cell will be pumped by an air operated diaphragm pump (one operating and one spare) to the moly third cleaner flotation column cell.

Moly fifth cleaner concentrate will flow by gravity to moly concentrate thickener, and tailings will be pumped by an air operated diaphragm pump (one operating/one spare) to the moly fourth cleaner column cell.

Flotation air for the moly cleaner flotation column cells will be supplied by the column cell compressed air system.

Flotation reagents will be added at several points in the moly flotation circuit.

Reagent addition points and quantities and slurry pH levels are generally as used for tests conducted at METCON.

17.6 Copper Concentrate Dewatering

Tailings (copper concentrate) from the moly separation (rougher) flotation and tailings (copper concentrate) from the moly first cleaner flotation circuits will be pumped to the copper concentrate thickener. Thickener overflow will flow by gravity to the copper concentrate thickener overflow surge tank, from where it will flow to the thickener overflow manifold. The combined copper concentrate thickener overflow and the bulk concentrate thickener overflow will flow by gravity to the process water pond. Copper thickener underflow will be pumped by a variable speed horizontal centrifugal slurry pump (one operating and one stand-by) to the copper concentrate filter feed tank.


SGS FORM No. A263a-2/14/13 !

! 149

Copper concentrate slurry will be pumped to the copper concentrate filter. Filter cake will discharge to the copper concentrate conveyor that will discharge to a covered concentrate load-out stockpile. Filtrate from the copper concentrate filter will be pumped back to copper concentrate thickener.

Copper concentrate will be reclaimed by front-end loader and loaded onto highway haulage trucks. A truck scale will be located at the plant entrance, which is near the concentrate load out area.

17.7 Molybdenite Concentrate Dewatering

Molybdenite concentrate from the molybdenite fifth cleaner flotation column cell will flow by gravity to moly concentrate thickener.

Thickener overflow will flow by gravity to the moly cleaner tails pump box, from where it will be pumped to the moly middling thickener. Moly thickener underflow will be pumped by a variable speed horizontal centrifugal slurry pump (one operating and one stand-by) to the moly concentrate filter feed tank. Concentrate slurry will be pumped from the moly filter feed tank by moly filter feed pump to the molybdenite concentrate filter.

Molybdenite concentrate slurry will be dewatered by moly concentrate filter. Filter filtrate will flow by gravity to the moly cleaner tails pump box, from where it will be pumped to the moly middling thickener. Filter cake will discharge to a moly concentrate dryer. The moly concentrate dryer will discharge to the molybdenite concentrate bin.

Molybdenite concentrate will be withdrawn from the molybdenite concentrate bin to the packaging system. Molybdenite concentrate will be bagged in Super-sacs for shipment by trucks to market.

17.8 Tailings

When the backfill plant is in operation tailings from the flotation circuit will be routed to a cyclone plant to recover that sand fraction for mine backfill. When the backfill plant is in operation, rejects from the backfill plant (slimes) will be thickened to approximately 50 percent solids by weight prior to transport to the TSF. When the backfill plant is not in operation, the flotation tailings discharge will bypass the cyclone plant and be transported to the thickener and TSF.


SGS FORM No. A263a-2/14/13 !

! 150

18.0 INFRASTRUCTURE

18.1 Access Roads/Project Roads

Currently there is an unpaved dirt road access to the proposed mine and mill site. For this study it was assumed that this 10 mile access road will be upgraded and paved to improve access to the plant site.

18.2 Power Supply and Distribution

Power supply options considered for the Redhawk mine include the local utility and on-site power generation, using natural gas as fuel. The local utility, APS (Arizona Public Service), is the primary option whereas, on-site generation using gas turbine engine for generating electricity was considered as a potential alternative.

The Redhawk mine is within the service territory of APS. Utility power is considered to be supplied to the mine via a high voltage overhead transmission line, about 20 miles long, from San Manuel substation. The line will be capable of handling the entire power requirement of 50MW for the Redhawk mine project. The power supply voltage will be stepped down to distribution voltage via overhead and/or underground power lines to the different facilities of the mine. The main substation will be located in close proximity to the concentrator that houses the biggest loads of the mine. The distribution voltage will be stepped down as required for medium voltage and low voltage applications. A SCADA (Supervisory Control And Data Acquisition) system will be utilized for controlling mechanical systems, based on feedback signals from motor control centers and instruments via PLCs (programmable logic controllers). All electrical installations, including transformers, switchgear, motor control centers, instrument controls, lighting and grounding design to be as per the National Electric Code (NEC) requirements and in accordance with NEMA and ANSI standards.

The power supply option for on-site power generation includes installing a gas pipeline for transporting high pressure natural gas to a gas turbine engine, installed at the mine site, which will generate electricity and be transmitted via the on-site power distribution system.

18.3 Water Supply and Distribution

The investigation of a source for mine water is in progress. The well field target has been selected in the alluvial pediment of Galiuro Mountains to the west of the mining area on state owned land. Ground-water withdrawal is prohibited inside a three mile buffer zone that parallels the San Pedro River. Well targets have been identified outside the buffer along Redhawk’s state land right of way, and a permit to drill exploration wells has been submitted to the Arizona Department of Water Resources. The proposed well field is located in the vicinity of an existing state-owned well where preliminary pumping tests yielded encouraging results.


SGS FORM No. A263a-2/14/13 !

! 151

The milling and process system requires raw water from water wells. The well field is located about 10.6 miles west of the site. The wells are connected via distribution pipelines to a holding tank at the first transfer pump station. Three pump stations are required to transport the water to the final holding tank above the plant site.

The raw water distribution system provides raw water only for process requirements such as reagent mixing, and gland water. The firewater system and potable water system do not draw water from the process raw water tank. Process raw water storage is sufficient for a short (nominally two to three hours) disruption of the raw water supply pumps. All other reagent preparation systems use water from the process raw water tank. Gland water pumps also draw directly from the raw water tank.

The copper-moly process water tank receives tailings thickener overflow and tailings reclaim water and raw water if sufficient reclaim water is not available. The water is pumped to the grinding circuit and may contain a small amount of solids so it is not suitable for general distribution throughout the process plant.

Water reclaimed from the copper thickener, moly thickener, copper filter and moly filter will contain residual hydrosulfide and will be recycled internal to the moly plant as operations allow.

18.4 Project Water Balance

A water balance for the Tailings impoundment is included in Table 18.1. The primary water loss is associated with water that is permanently bound or entrained within the pore space of the tailings solids. Additional losses include evaporation from the TSF pond free water surface and water lost to evaporation on the tailings beach.

Post deposition dry densities of 60 and 68 pounds per cubic foot (pcf) are used to estimate the entrained water losses for the backfill plant rejects and whole tailings (backfill plant by-pass), respectively. Impoundment volume assumes an average post deposition dry density of 75 pcf. The higher post deposition density reflects ongoing consolidation of the tailings through both natural processes and managed deposition.


SGS FORM No. A263a-2/14/13 !


Table 18.1 WATER BALANCE CALCULATIONS Backfill Plant Rejects Thickened

Date: 5/30/2013

Project No.: 123-92517 Date

Subject: Backfill Plant Rejects Thickened Made by: DMW 5/23/2

013

Project Short Title:

Copper Creek Preliminary Economic Assessment Checked: GM 5/24/2

013

Reviewed: MJG

Precipitation and Evaporation Data

2 3 4 5 6 7 8 9 10 Variables used in Water Balance

Month

Days

Precip., in

Pan evap., in

Storms/month

Precip/storm, in

S

Runoff/storm, in

Monthly runoff, in

Average water input

2,261 gpm

Beach Loss 50% of Pan

Evap

Average Entrained Water

1,601 gpm

January

31 1.77 2.63 2 0.8

9 1.24 0.22 0.43 Max Pond Area

2,178,

000

50 acre

February

28 1.81 3.70 2 0.9

0 1.24 0.23 0.45 Total TSF catchment area

41,284,950 ft^2

March

31 1.88 5.63 2 0.9

4 1.24 0.25 0.50

April

30 0.78 8.00 1 0.7

8 1.24 0.16 0.16 TSF Free Water area 20%

of Tailings Area

May 31 0.3

8 10.42 1 0.38 1.24 0.01 0.01 TSF Free Water Pan

Evap. coefficient 75%

Document No.

Q400-05-028 25 July 2013

SGS FORM No. A263a-2/14/13 !


June

30 0.33 12.08 1 0.3

3 1.24 0.01 0.01 Reclaim Pump capacity

2,500 gpm

July 31 2.6

9 10.67 3 0.90 1.24 0.22 0.67 Minimum

Water Storage 100,0

00 gal

August

31 3.22 9.44 3 1.0

7 1.24 0.33 0.99 CN-Native ground areas 89

September

30 1.58 8.38 2 0.7

9 1.24 0.17 0.33

October

31 1.34 6.33 2 0.6

7 1.24 0.11 0.22

November

30 1.27 3.83 2 0.6

4 1.24 0.09 0.19

December

31 1.94 2.73 2 0.9

7 1.24 0.27 0.53

Delivered

Post Deposition

Delivered

Surface Area 3rd Order Polynomial Expression Input

Tons per day

Dry weight, pcf

V solids, ft3

Vwater, ft3

Water In, gpd

Whole Tailings

2947.0 68 0.41 0.59 1,038,

889

Coefficient

Degree

Backfill Rejects

6393.0 60 0.36 0.64 2,217,

628

0 x5 Total 9340.0 3,256,

517

0 x4

5160 x3

Entrained water, not backfilling

1,466,322.65

gpdthickener UF

-163920

x2Entrained

water, backfilling

839,722.22 gpd

thickener UF

Document No.

Q400-05-028 25 July 2013

SGS FORM No. A263a-2/14/13 !


3000000

x

454652

constant

Results

Average water in 2,261 gpm

Average tailings reclaim

462 gpm

Notes:

Average makeup for tailings loss

1,800 gpm

Tailings direct precipitation reports 100% to water balance

Maximum make-up rate 2261 gpm

Assume maximum pond size 50 acres

Document No.

Q400-05-028 25 July 2013

SGS FORM No. A263a-2/14/13 !


Year

Month

Elapsed

Years

Impoundm

ent Tailin

gs Area, ft^2

Pond Area, ft^2

Beach Area,

ft^2

Water Inflows Water Losses

Monthly Balance at

TSF, gal

Water Reclaimed rom TSF, gal

End of

Month

Water Storage, gal

Water Lost to

Tailings,

gpm

Tailings

Water, gal

Direct

Precip., gal

Runon f/

Undiv.Area, gal

Total Entrained Water, gal

Impoundment Free

Water Evap., gal

Impoundment

BeachEvap., gal

1

Jan-15

0.08

708,267

141,65

3

566,61

4

100,952,03

0- 10,99

4,83671,487,391

174,344.12

464,918

39,820,2

14

39,720,214

100,000 1,372

Feb-15

0.16

935,322

187,06

4

748,25

8

91,182,479

1,052,44

5

11,423,088

64,569,256

323,745.87

863,322

38,001,6

8737,901,687

100,000 1,321

Mar-15

0.25

1,184,489

236,89

8

947,59

1

100,952,03

0

1,389,66

2

12,471,638

71,487,391

623,508.89

1,662,69

0

41,139,7

4041,039,740

100,000 1,342

Apr-1

50.33

1,423,419

284,68

4

1,138,

735

97,695,513

690,764

3,968,871

69,181,346

1,065,163.9

5

2,840,43

7

29,368,2

0129,268,201

100,000 1,584

May-15

0.41

1,668,059

333,61

2

1,334,

447

100,952,03

0

391,399

302,048

71,487,391

1,625,985.5

8

4,335,96

2

24,296,1

3924,196,139

100,000 1,719

Jun-15

0.50

1,902,643

380,52

9

1,522,1

15

97,695,513

395,710

138,722

69,181,346

2,148,630.7

6

5,729,68

2

21,270,2

8621,170,286

100,000 1,771

Jul-15

0.58

2,142,830

428,56

6

1,714,

264

100,952,03

0

3,599,59

6

16,442,772

71,487,391

2,137,470.8

5

5,699,92

2

41,769,6

1441,669,614

100,000 1,328

Aug-15

0.67

2,380,781

476,15

6

1,904,

625

100,952,03

0

4,773,78

5

24,027,366

71,487,391

2,102,542.4

1

5,606,78

0

50,656,4

6750,556,467

100,000 1,129

Sep-15

0.75

2,608,946

521,78

9

2,087,

157

97,695,513

2,573,90

4

8,019,958

69,181,346

2,045,183.2

8

5,453,82

2

31,709,0

2331,609,023

100,000 1,530

Oct-1

50.83

2,842,555

568,51

1

2,274,

044

100,952,03

0

2,379,92

9

5,198,253

71,487,391

1,681,460.3

5

4,483,89

4

30,977,4

6630,877,466

100,000 1,570

Nov-

150.92

3,066,553

613,31

1

2,453,

242

97,695,513

2,436,64

5

4,460,024

69,181,346

1,098,596.8

4

2,929,59

2

31,482,6

4831,382,648

100,000 1,535

Dec-

15

1.00

3,295,892

659,17

8

2,636,

714

100,952,03

0

3,989,83

0

12,660,273

71,487,391

841,258.83

2,243,35

7

43,130,1

2643,030,126

100,000 1,298

Document No.

Q400-05-028 25 July 2013

SGS FORM No. A263a-2/14/13 !


2

Jan-16

1.08

3,523,090

704,61

8

2,818,

472

100,952,03

0

3,889,10

4

10,232,120

71,487,391

867,229.07

2,312,61

1

40,506,0

23

40,406,023

100,000 1,356

Feb-16

1.16

3,726,476

745,29

5

2,981,

181

91,182,479

4,193,11

0

10,632,905

64,569,256

1,289,856.0

8

3,439,61

6

36,809,7

6536,709,765

100,000 1,351

Mar-16

1.25

3,949,650

789,93

0

3,159,

720

100,952,03

0

4,633,79

5

11,611,646

71,487,391

2,079,075.1

3

5,544,20

0

38,186,8

0538,086,805

100,000 1,408

Apr-1

61.33

4,163,638

832,72

8

3,330,911

97,695,513

2,020,55

2

3,696,037

69,181,346

3,115,707.7

9

8,308,55

4

22,906,4

9422,806,494

100,000 1,734

May-16

1.41

4,382,725

876,54

5

3,506,

180

100,952,03

0

1,028,37

7

281,351

71,487,391

4,272,180.2

1

11,392,4

81

15,209,7

0615,109,706

100,000 1,923

Jun-16

1.50

4,592,794

918,55

9

3,674,

235

97,695,513

955,206

129,246

69,181,346

5,186,583.0

8

13,830,8

88

10,681,1

4710,581,147

100,000 2,017

Jul-16

1.58

4,807,866

961,57

3

3,846,

293

100,952,03

0

8,076,41

2

15,323,247

71,487,391

4,795,842.7

2

12,788,9

14

35,379,5

4235,279,542

100,000 1,471

Aug-16

1.67

5,020,927

1,004,

185

4,016,

742

100,952,03

0

10,067,6

32

22,396,801

71,487,391

4,434,138.5

4

11,824,3

69

45,770,5

6445,670,564

100,000 1,238

Sep-16

1.75

5,225,218

1,045,

044

4,180,

174

97,695,513

5,155,03

5

7,477,441

69,181,346

4,096,109.1

1

10,922,9

58

26,227,5

7626,127,576

100,000 1,657

Oct-1

61.83

5,434,376

1,086,

875

4,347,

500

100,952,03

0

4,549,93

1

4,847,782

71,487,391

3,214,603.7

4

8,572,27

7

27,175,4

7227,075,472

100,000 1,655

Nov-

161.92

5,634,924

1,126,

985

4,507,

939

97,695,513

4,477,44

1

4,160,300

69,181,346

2,018,719.5

0

5,383,25

2

29,849,9

3629,749,936

100,000 1,573

Dec-

16

2.00

5,840,252

1,168,

050

4,672,

202

100,952,03

0

7,069,89

5

11,812,337

71,487,391

1,490,693.1

3

3,975,18

2

42,980,9

9642,880,996

100,000 1,301

Year

Month

Elapsed

Years

Impoundm

ent Tailin

gs Area, ft^2

Pond Area, ft^2

Beach Area,

ft^2


Monthly Balance at

TSF, gal


End of

Month

Water Storage, gal

Water Lost to

Tailings,

gpm

Tailings

Water, gal

Direct

Precip., gal

Runon f/

Undiv.Area, gal


Impoundment Free

Water Evap., gal

Impoundment

BeachEvap., gal

Document No.

Q400-05-028 25 July 2013

SGS FORM No. A263a-2/14/13 !


3

Jan-17

2.08

6,043,662

1,208,

732

4,834,

929

100,952,03

0

6,671,53

9

9,549,135

71,487,391

1,487,682.5

4

3,967,15

3

40,330,4

76

40,230,476

100,000 1,360

Feb-17

2.16

6,225,754

1,245,

151

4,980,

603

91,182,479

7,005,35

1

9,925,353

64,569,256

2,154,938.5

4

5,746,50

3

35,742,4

8535,642,485

100,000 1,377

Mar-17

2.25

6,425,565

1,285,113

5,140,

452

100,952,03

0

7,538,58

0

10,841,612

71,487,391

3,382,383.7

8

9,019,69

0

35,542,7

5735,442,757

100,000 1,468

Apr-1

72.33

6,617,157

1,323,

431

5,293,

725

97,695,513

3,211,20

7

3,451,749

69,181,346

4,951,709.0

7

13,204,5

58

17,120,8

5717,020,857

100,000 1,867

May-17

2.41

6,813,320

1,362,

664

5,450,

656

100,952,03

0

1,598,70

0

262,820

71,487,391

6,641,468.0

7

17,710,5

82

7,074,10

96,974,

109100,0

00 2,105

Jun-17

2.50

7,001,417

1,400,

283

5,601,

133

97,695,513

1,456,15

0

120,762

69,181,346

7,906,610.3

0

21,084,2

94

1,200,17

41,100,

174100,0

00 2,236

Jul-17

2.58

7,194,005

1,438,

801

5,755,

204

100,952,03

0

12,084,7

27

14,320,881

71,487,391

7,176,014.3

1

19,136,0

38

29,658,1

9429,558,194

100,000 1,599

Aug-17

2.67

7,384,806

1,476,

961

5,907,

844

100,952,03

0

14,807,5

24

20,936,861

71,487,391

6,521,753.4

9

17,391,3

43

41,395,9

2741,295,927

100,000 1,336

Sep-17

2.75

7,567,766

1,513,

553

6,054,

213

97,695,513

7,466,12

0

6,991,684

69,181,346

5,932,459.8

0

15,819,8

93

21,319,6

1721,219,617

100,000 1,770

Oct-1

72.83

7,755,102

1,551,

020

6,204,

082

100,952,03

0

6,492,96

0

4,533,969

71,487,391

4,587,386.4

4

12,233,0

31

23,771,1

5123,671,151

100,000 1,731

Nov-

172.92

7,934,746

1,586,

949

6,347,

797

97,695,513

6,304,85

1

3,891,914

69,181,346

2,842,633.9

3

7,580,35

7

28,387,9

4128,287,941

100,000 1,607

Dec-

17

3.00

8,118,692

1,623,

738

6,494,

954

100,952,03

0

9,828,05

2

11,053,022

71,487,391

2,072,252.7

7

5,526,00

7

42,847,4

5342,747,453

100,000 1,304

Year

Month

Elapsed

Years

Impoundm

ent Tailin

gs Area, ft^2

Pond Area, ft^2

Beach Area,

ft^2


Monthly Balance at

TSF, gal


End of

Month

Water Storage, gal

Water Lost to

Tailings,

gpm

Tailings

Water, gal

Direct

Precip., gal

Runon f/

Undiv.Area, gal


Impoundment Free

Water Evap., gal

Impoundment

BeachEvap., gal

Document No.

Q400-05-028 25 July 2013

SGS FORM No. A263a-2/14/13 !


4

Jan-18

3.08

8,300,943

1,660,

189

6,640,

754

100,952,03

0

9,163,33

0

8,937,492

71,487,391

2,043,325.5

2

5,448,86

8

40,173,2

67

40,073,267

100,000 1,364

Feb-18

3.16

8,464,116

1,692,

823

6,771,

293

91,182,479

9,524,00

4

9,291,666

64,569,256

2,929,709.5

6

7,812,55

9

34,786,6

2534,686,625

100,000 1,401

Mar-18

3.25

8,643,194

1,728,

639

6,914,

555

100,952,03

0

10,140,3

39

10,151,908

71,487,391

4,549,732.0

1

12,132,6

19

33,174,5

3533,074,535

100,000 1,521

Apr-18

3.33

8,814,933

1,762,

987

7,051,

947

97,695,513

4,277,75

5

3,232,924

69,181,346

6,596,335.5

7

17,590,2

28

11,938,2

8311,838,283

100,000 1,987

May-18

3.41

8,990,803

1,798,

161

7,192,

642

100,952,03

0

2,109,63

2

246,218

71,487,391

8,764,028.2

6

23,370,7

42- (100,0

00)100,0

00 2,264

Jun-18

3.50

9,159,473

1,831,

895

7,327,

578

97,695,513

1,904,98

1

113,160

69,181,346

10,343,675.

15

27,583,1

34- (100,0

00)100,0

00 2,264

Jul-18

3.58

9,332,207

1,866,

441

7,465,

765

100,952,03

0

15,676,5

48

13,422,668

71,487,391

9,308,868.2

1

24,823,6

49

24,531,3

3824,431,338

100,000 1,714

Aug-18

3.67

9,503,375

1,900,

675

7,602,

700

100,952,03

0

19,055,5

40

19,628,424

71,487,391

8,392,729.0

1

22,380,6

11

37,475,2

6337,375,263

100,000 1,424

Sep-18

3.75

9,667,550

1,933,

510

7,734,

040

97,695,513

9,537,70

1

6,556,267

69,181,346

7,578,505.2

7

20,209,3

47

16,920,2

8216,820,282

100,000 1,872

Oct-18

3.83

9,835,695

1,967,

139

7,868,

556

100,952,03

0

8,234,93

6

4,252,627

71,487,391

5,818,122.2

6

15,514,9

93

20,719,0

8720,619,087

100,000 1,800

Nov-18

3.92

9,996,979

1,999,

396

7,997,

583

97,695,513

7,943,47

5

3,651,255

69,181,346

3,581,431.6

0

9,550,48

4

27,076,9

8126,976,981

100,000 1,637

Dec-18

4.00

10,162,172

2,032,

434

8,129,

738

100,952,03

0

12,301,7

79

10,372,009

71,487,391

2,593,840.1

3

6,916,90

7

42,727,6

8042,627,680

100,000 1,307

Year

Month

Elapsed

Years

Impoundm

ent Tailin

gs Area, ft^2

Pond Area, ft^2

Beach Area,

ft^2


Monthly Balance at

TSF, gal


End of

Month

Water Storage, gal

Water Lost to

Tailings,

gpm

Tailings

Water, gal

Direct

Precip., gal

Runon f/

Undiv.Area, gal


Impoundment Free

Water Evap., gal

Impoundment

BeachEvap., gal

Document No.

Q400-05-028 25 July 2013

SGS FORM No. A263a-2/14/13 !


5

Jan-19

4.08

10,325,894

2,065,

179

8,260,

715

100,952,03

0

11,398,6

53

8,388,802

71,487,391

2,541,778.9

8

6,778,07

7

40,032,2

38

39,932,238

100,000 1,367

Feb-19

4.16

10,472,523

2,094,

505

8,378,

018

91,182,479

11,783,9

07

8,723,081

64,569,256

3,624,885.3

8

9,666,36

1

33,928,9

6433,828,964

100,000 1,422

Mar-19

4.25

10,633,497

2,126,

699

8,506,

797

100,952,03

0

12,475,3

95

9,532,904

71,487,391

5,597,417.0

0

14,926,4

45

31,049,0

7530,949,075

100,000 1,568

Apr-19

4.33

10,787,929

2,157,

586

8,630,

343

97,695,513

5,235,22

1

3,036,480

69,181,346

8,072,755.0

8

21,527,3

47

7,285,76

67,185,

766100,0

00 2,095

May-19

4.41

10,946,133

2,178,

000

8,768,

133

100,952,03

0

2,568,43

7

231,310

71,487,391

10,615,322.

40

28,489,9

18- (100,0

00)100,0

00 2,264

Jun-19

4.50

11,097,921

2,178,

000

8,919,

921

97,695,513

2,308,13

8

106,332

69,181,346

12,297,937.

61

33,577,1

76- (100,0

00)100,0

00 2,264

Jul-19

4.58

11,253,430

2,178,

000

9,075,

430

100,952,03

0

18,903,8

82

12,615,603

71,487,391

10,862,765.

82

30,175,7

79

20,045,5

7919,945,579

100,000 1,815

Aug-19

4.67

11,407,597

2,178,

000

9,229,

597

100,952,03

0

22,873,7

59

18,452,369

71,487,391

9,617,300.6

4

27,169,8

22

34,103,6

4434,003,644

100,000 1,500

Sep-19

4.75

11,555,531

2,178,

000

9,377,

531

97,695,513

11,400,3

23

6,164,770

69,181,346

8,536,797.7

4

24,503,8

53

13,138,6

1013,038,610

100,000 1,960

Oct-19

4.83

11,707,113

2,178,

000

9,529,1

13

100,952,03

0

9,801,78

1

3,999,571

71,487,391

6,441,776.8

4

18,789,2

32

18,134,9

8218,034,982

100,000 1,857

Nov-19

4.92

11,852,581

2,178,

000

9,674,

581

97,695,513

9,417,91

5

3,434,709

69,181,346

3,901,357.8

1

11,553,1

08

26,012,3

2525,912,325

100,000 1,662

Dec-19

5.00

12,001,652

2,178,

000

9,823,

652

100,952,03

0

14,528,5

55

9,758,982

71,487,391

2,779,614.3

4

8,358,11

6

42,714,4

4642,614,446

100,000 1,307

Year

Month

Elapsed

Years

Impoundm

ent Tailin

gs Area, ft^2

Pond Area, ft^2

Beach Area,

ft^2


Monthly Balance at

TSF, gal


End of

Month

Water Storage, gal

Water Lost to

Tailings,

gpm

Tailings

Water, gal

Direct

Precip., gal

Runon f/

Undiv.Area, gal


Impoundment Free

Water Evap., gal

Impoundment

BeachEvap., gal

Document No.

Q400-05-028 25 July 2013

SGS FORM No. A263a-2/14/13 !


6

Jan-20

5.08

12,149,474

2,178,

000

9,971,

474

100,952,03

0

13,411,6

85

7,894,677

71,487,391

2,680,636.9

7

8,181,78

8

40,008,5

76

39,908,576

100,000 1,367

Feb-20

5.16

12,281,934

2,178,

000

10,103,93

4

91,182,479

13,819,8

95

8,210,832

64,569,256

3,769,387.9

9

11,657,6

82

33,316,8

7933,216,879

100,000 1,438

Mar-20

5.25

12,427,433

2,178,

000

10,249,43

3

100,952,03

0

14,580,0

71

8,974,972

71,487,391

5,732,439.0

6

17,984,1

60

29,403,0

8329,303,083

100,000 1,605

Apr-20

5.33

12,567,103

2,178,

000

10,389,10

3

97,695,513

6,098,62

7

2,859,334

69,181,346

8,149,136.3

9

25,914,3

60

3,508,63

13,408,

631100,0

00 2,183

May-20

5.41

12,710,272

2,178,

000

10,532,27

2

100,952,03

0

2,982,38

1

217,860

71,487,391

10,615,322.

40

34,222,0

59- (100,0

00)100,0

00 2,264

Jun-20

5.50

12,847,723

2,178,

000

10,669,72

3

97,695,513

2,672,06

0

100,168

69,181,346

12,297,937.

61

40,163,9

37- (100,0

00)100,0

00 2,264

Jul-20

5.58

12,988,636

2,178,

000

10,810,63

6

100,952,03

0

21,818,7

38

11,886,680

71,487,391

10,862,765.

82

35,945,3

33

16,461,9

5716,361,957

100,000 1,895

Aug-20

5.67

13,128,430

2,178,

000

10,950,43

0

100,952,03

0

26,324,2

60

17,389,576

71,487,391

9,617,300.6

4

32,235,5

62

31,425,6

1231,325,612

100,000 1,560

Sep-20

5.75

13,262,668

2,178,

000

11,084,

668

97,695,513

13,084,5

31

5,810,774

69,181,346

8,536,797.7

4

28,964,6

69

10,008,0

059,908,

005100,0

00 2,032

Oct-20

5.83

13,400,318

2,178,

000

11,222,

318

100,952,03

0

11,219,4

17

3,770,612

71,487,391

6,441,776.8

4

22,127,8

44

15,985,0

4715,885,047

100,000 1,906

Nov-20

5.92

13,532,515

2,178,

000

11,354,

515

97,695,513

10,752,7

69

3,238,664

69,181,346

3,901,357.8

1

13,559,2

36

25,145,0

0525,045,005

100,000 1,682

Year

Month

Elapsed

Years

Impoundm

ent Tailin

gs Area, ft^2

Pond Area, ft^2

Beach Area,

ft^2


Monthly Balance at

TSF, gal


End of

Month

Water Storage, gal

Water Lost to

Tailings,

gpm

Tailings

Water, gal

Direct

Precip., gal

Runon f/

Undiv.Area, gal


Impoundment Free

Water Evap., gal

Impoundment

BeachEvap., gal

Document No.

Q400-05-028 25 July 2013

SGS FORM No. A263a-2/14/13 !


Dec-20

6.00

13,668,092

2,178,

000

11,490,

092

100,952,03

0

16,545,8

57

9,203,623

71,487,391

2,779,614.3

4

9,775,94

9

42,758,5

5642,658,556

100,000 1,306

7

Jan-21

6.08

13,802,644

2,178,

000

11,624,

644

100,952,03

0

15,236,6

03

7,446,727

71,487,391

2,680,636.9

7

9,538,24

6

40,029,0

85

39,929,085

100,000 1,367

Feb-21

6.16

13,923,310

2,178,

000

11,745,

310

91,182,479

15,666,8

06

7,746,154

64,569,256

3,769,387.9

9

13,551,4

63

32,805,3

3132,705,331

100,000 1,450

Mar-21

6.25

14,055,964

2,178,

000

11,877,

964

100,952,03

0

16,490,6

91

8,468,483

71,487,391

5,732,439.0

6

20,841,6

61

27,949,7

1327,849,713

100,000 1,638

Apr-21

6.33

14,183,415

2,178,

000

12,005,41

5

97,695,513

6,883,00

0

2,698,404

69,181,346

8,149,136.3

9

29,946,0

57

100,377 377 100,0

00 2,261

May-21

6.41

14,314,179

2,178,

000

12,136,17

9

100,952,03

0

3,358,72

7

205,631

71,487,391

10,615,322.

40

39,433,5

64- (100,0

00)100,0

00 2,264

Jun-21

6.50

14,439,837

2,178,

000

12,261,83

7

97,695,513

3,003,18

7

94,560

69,181,346

12,297,937.

61

46,157,1

18- (100,0

00)100,0

00 2,264

Jul-21

6.58

14,568,784

2,178,

000

12,390,78

4

100,952,03

0

24,473,1

22

11,222,893

71,487,391

10,862,765.

82

41,199,3

21

13,198,5

6813,098,568

100,000 1,968

Aug-21

6.67

14,696,835

2,178,

000

12,518,83

5

100,952,03

0

29,469,1

22

16,420,923

71,487,391

9,617,300.6

4

36,852,5

88

28,984,7

9628,884,796

100,000 1,614

Sep-21

6.75

14,819,922

2,178,

000

12,641,92

2

97,695,513

14,620,8

69

5,487,858

69,181,346

8,536,797.7

4

33,033,8

33

7,152,26

27,052,

262100,0

00 2,098

Oct-21

6.83

14,946,268

2,178,

000

12,768,26

8

100,952,03

0

12,513,7

63

3,561,566

71,487,391

6,441,776.8

4

25,176,1

04

14,022,0

8713,922,087

100,000 1,950

Year

Month

Elapsed

Years

Impoundm

ent Tailin

gs Area, ft^2

Pond Area, ft^2

Beach Area,

ft^2


Monthly Balance at

TSF, gal


End of

Month

Water Storage, gal

Water Lost to

Tailings,

gpm

Tailings

Water, gal

Direct

Precip., gal

Runon f/

Undiv.Area, gal


Impoundment Free

Water Evap., gal

Impoundment

BeachEvap., gal

Document No.

Q400-05-028 25 July 2013

SGS FORM No. A263a-2/14/13 !


Nov-21

6.92

15,067,739

2,178,

000

12,889,73

9

97,695,513

11,972,6

39

3,059,506

69,181,346

3,901,357.8

1

15,392,5

57

24,352,3

9624,252,396

100,000 1,700

Dec-21

7.00

15,192,452

2,178,

000

13,014,45

2

100,952,03

0

18,391,1

66

8,695,613

71,487,391

2,779,614.3

4

11,072,8

98

42,798,9

0542,698,905

100,000 1,305

8

Jan-22

7.08

15,316,364

2,178,

000

13,138,36

4

100,952,03

0

16,907,5

83

7,036,562

71,487,391

2,680,636.9

7

10,780,2

83

40,047,8

65

39,947,865

100,000 1,367

Feb-22

7.16

15,427,610

2,178,

000

13,249,61

0

91,182,479

17,359,4

77

7,320,282

64,569,256

3,769,387.9

9

15,287,0

90

32,336,5

0432,236,504

100,000 1,462

Mar-22

7.25

15,550,049

2,178,

000

13,372,04

9

100,952,03

0

18,243,5

76

8,003,808

71,487,391

5,732,439.0

6

23,463,2

55

26,616,3

2826,516,328

100,000 1,667

Apr-22

7.33

15,667,827

2,178,

000

13,489,82

7

97,695,513

7,603,36

2

2,550,606

69,181,346

8,149,136.3

9

33,648,7

41- (100,0

00)100,0

00 2,264

May-22

7.41

15,788,814

2,178,

000

13,610,81

4

100,952,03

0

3,704,74

0

194,388

71,487,391

10,615,322.

40

44,225,0

33- (100,0

00)100,0

00 2,264

Jun-22

7.50

15,905,224

2,178,

000

13,727,22

4

97,695,513

3,307,95

7

89,398

69,181,346

12,297,937.

61

51,673,2

60- (100,0

00)100,0

00 2,264

Jul-22

7.58

16,024,834

2,178,

000

13,846,83

4

100,952,03

0

26,919,0

43

10,611,238

71,487,391

10,862,765.

82

46,040,6

84

10,191,4

7010,091,470

100,000 2,035

Aug-22

7.67

16,143,772

2,178,

000

13,965,77

2

100,952,03

0

32,370,4

24

15,527,289

71,487,391

9,617,300.6

4

41,112,0

39

26,733,0

1326,633,013

100,000 1,665

Year

Month

Elapsed

Years

Impoundm

ent Tailin

gs Area, ft^2

Pond Area, ft^2

Beach Area,

ft^2


Monthly Balance at

TSF, gal


End of

Month

Water Storage, gal

Water Lost to

Tailings,

gpm

Tailings

Water, gal

Direct

Precip., gal

Runon f/

Undiv.Area, gal


Impoundment Free

Water Evap., gal

Impoundment

BeachEvap., gal

Document No.

Q400-05-028 25 July 2013

SGS FORM No. A263a-2/14/13 !


Sep-22

7.75

16,258,252

2,178,

000

14,080,25

2

97,695,513

16,039,8

80

5,189,602

69,181,346

8,536,797.7

4

36,792,2

45

4,514,60

54,414,

605100,0

00 2,159

Oct-22

7.83

16,375,924

2,178,

000

14,197,92

4

100,952,03

0

13,710,7

43

3,368,245

71,487,391

6,441,776.8

4

27,995,0

59

12,206,7

9212,106,792

100,000 1,990

Nov-22

7.92

16,489,213

2,178,

000

14,311,213

97,695,513

13,102,1

25

2,893,622

69,181,346

3,901,357.8

1

17,090,0

41

23,618,5

1523,518,515

100,000 1,717

Dec-22

8.00

16,605,692

2,178,

000

14,427,69

2

100,952,03

0

20,101,9

58

8,224,635

71,487,391

2,779,614.3

4

12,275,3

04

42,836,3

1342,736,313

100,000 1,304

9

Jan-23

8.08

16,721,592

2,178,

000

14,543,59

2

100,952,03

0

18,458,8

02

6,655,795

71,487,391

2,680,636.9

7

11,933,3

00

40,065,2

99

39,965,299

100,000 1,366

Feb-23

8.16

16,825,795

2,178,

000

14,647,79

5

91,182,479

18,932,7

45

6,924,453

64,569,256

3,769,387.9

9

16,900,2

83

31,900,7

4931,800,749

100,000 1,473

Mar-23

8.25

16,940,648

2,178,

000

14,762,64

8

100,952,03

0

19,875,0

49

7,571,318

71,487,391

5,732,439.0

6

25,903,2

68

25,375,2

9925,275,299

100,000 1,695

Apr-23

8.33

17,051,297

2,178,

000

14,873,29

7

97,695,513

8,274,73

9

2,412,859

69,181,346

8,149,136.3

9

37,099,6

40- (100,0

00)100,0

00 2,264

May-23

8.41

17,165,138

2,178,

000

14,987,13

8

100,952,03

0

4,027,68

5

183,895

71,487,391

10,615,322.

40

48,697,0

60- (100,0

00)100,0

00 2,264

Jun-23

8.50

17,274,843

2,178,

000

15,096,84

3

97,695,513

3,592,81

0

84,574

69,181,346

12,297,937.

61

56,828,9

07- (100,0

00)100,0

00 2,264

Year

Month

Elapsed

Years

Impoundm

ent Tailin

gs Area, ft^2

Pond Area, ft^2

Beach Area,

ft^2


Monthly Balance at

TSF, gal


End of

Month

Water Storage, gal

Water Lost to

Tailings,

gpm

Tailings

Water, gal

Direct

Precip., gal

Runon f/

Undiv.Area, gal


Impoundment Free

Water Evap., gal

Impoundment

BeachEvap., gal

Document No.

Q400-05-028 25 July 2013

SGS FORM No. A263a-2/14/13 !


9Jul-23

8.58

17,387,746

2,178,

000

15,209,74

6

100,952,03

0

29,208,5

08

10,038,707

71,487,391

10,862,765.

82

50,572,3

64

7,376,72

57,276,

725100,0

00 2,098

Aug-23

8.67

17,500,201

2,178,

000

15,322,20

1

100,952,03

0

35,090,2

45

14,689,553

71,487,391

9,617,300.6

4

45,105,0

54

24,622,0

8324,522,083

100,000 1,712

Sep-23

8.75

17,608,618

2,178,

000

15,430,61

8

97,695,513

17,372,1

09

4,909,586

69,181,346

8,536,797.7

4

40,320,8

05

2,038,25

91,938,

259100,0

00 2,217

Oct-23

8.83

17,720,246

2,178,

000

15,542,24

6

100,952,03

0

14,836,2

77

3,186,464

71,487,391

6,441,776.8

4

30,645,7

54

10,499,8

4810,399,848

100,000 2,028

Nov-23

8.92

17,827,898

2,178,

000

15,649,89

8

97,695,513

14,165,8

27

2,737,400

69,181,346

3,901,357.8

1

18,688,6

60

22,927,3

7522,827,375

100,000 1,733

Dec-23

9.00

17,938,772

2,178,

000

15,760,77

2

100,952,03

0

21,715,7

13

7,780,372

71,487,391

2,779,614.3

4

13,409,5

10

42,871,6

0042,771,600

100,000 1,303

Jan-24

9.08

18,049,291

2,178,

000

15,871,29

1

100,952,03

0

19,924,4

35

6,296,036

71,487,391

2,680,636.9

7

13,022,7

02

40,081,7

71

39,981,771

100,000 1,366

Feb-24

9.16

18,148,825

2,178,

000

15,970,82

5

91,182,479

20,421,4

46

6,549,899

64,569,256

3,769,387.9

9

18,426,7

63

31,488,4

1631,388,416

100,000 1,483

Mar-24

9.25

18,258,720

2,178,

000

16,080,72

0

100,952,03

0

21,421,4

34

7,161,384

71,487,391

5,732,439.0

6

28,216,0

24

24,198,9

9524,098,995

100,000 1,722

Apr-24

9.33

18,364,785

2,178,

000

16,186,78

5

97,695,513

8,912,15

6

2,282,080

69,181,346

8,149,136.3

9

40,375,9

78- (100,0

00)100,0

00 2,264

Year

Month

Elapsed

Years

Impoundm

ent Tailin

gs Area, ft^2

Pond Area, ft^2

Beach Area,

ft^2


Monthly Balance at

TSF, gal


End of

Month

Water Storage, gal

Water Lost to

Tailings,

gpm

Tailings

Water, gal

Direct

Precip., gal

Runon f/

Undiv.Area, gal


Impoundment Free

Water Evap., gal

Impoundment

BeachEvap., gal

Document No.

Q400-05-028 25 July 2013

SGS FORM No. A263a-2/14/13 !


10

May-24

9.41

18,474,109

2,178,

000

16,296,10

9

100,952,03

0

4,334,82

6

173,915

71,487,391

10,615,322.

40

52,950,2

45- (100,0

00)100,0

00 2,264

Jun-24

9.50

18,579,656

2,178,

000

16,401,65

6

97,695,513

3,864,18

4

79,978

69,181,346

12,297,937.

61

61,740,6

00- (100,0

00)100,0

00 2,264

Jul-24

9.58

18,688,481

2,178,

000

16,510,48

1

100,952,03

0

31,393,5

25

9,492,296

71,487,391

10,862,765.

82

54,897,3

03

4,690,39

14,590,

391100,0

00 2,159

Aug-24

9.67

18,797,081

2,178,

000

16,619,08

1

100,952,03

0

37,690,6

64

13,888,595

71,487,391

9,617,300.6

4

48,922,7

73

22,603,8

2422,503,824

100,000 1,757

Sep-24

9.75

18,901,981

2,178,

000

16,723,98

1

97,695,513

18,648,1

00

4,641,391

69,181,346

8,536,797.7

4

43,700,4

12- (100,0

00)100,0

00 2,264

Oct-24

9.83

19,010,193

2,178,

000

16,832,19

3

100,952,03

0

15,916,2

85

3,012,034

71,487,391

6,441,776.8

4

33,189,2

36

8,861,94

68,761,

946100,0

00 2,065

Nov-24

9.92

19,114,754

2,178,

000

16,936,75

4

97,695,513

15,188,3

47

2,587,226

69,181,346

3,901,357.8

1

20,225,3

86

22,262,9

9522,162,995

100,000 1,748

Dec-24

10.00

19,222,652

2,178,

000

17,044,65

2

100,952,03

0

23,269,9

09

7,352,504

71,487,391

2,779,614.3

4

14,501,8

55

42,905,5

8442,805,584

100,000 1,303

Jan-25

10.08

19,330,418

2,178,

000

17,152,41

8

100,952,03

0

21,338,6

59

5,948,896

71,487,391

2,680,636.9

7

14,073,8

93

40,097,6

65

39,997,665

100,000 1,365

Feb-25

10.16

19,427,659

2,178,

000

17,249,65

9

91,182,479

21,860,4

17

6,187,858

64,569,256

3,769,387.9

9

19,902,2

51

31,089,8

5830,989,858

100,000 1,493

Year

Month

Elapsed

Years

Impoundm

ent Tailin

gs Area, ft^2

Pond Area, ft^2

Beach Area,

ft^2


Monthly Balance at

TSF, gal


End of

Month

Water Storage, gal

Water Lost to

Tailings,

gpm

Tailings

Water, gal

Direct

Precip., gal

Runon f/

Undiv.Area, gal


Impoundment Free

Water Evap., gal

Impoundment

BeachEvap., gal

Document No.

Q400-05-028 25 July 2013

SGS FORM No. A263a-2/14/13 !


11

Mar-25

10.25

19,535,227

2,178,

000

17,357,22

7

100,952,03

0

22,919,0

53

6,764,378

71,487,391

5,732,439.0

6

30,455,8

45

23,059,7

8622,959,786

100,000 1,747

Apr-25

10.33

19,639,252

2,178,

000

17,461,25

2

97,695,513

9,530,63

6

2,155,185

69,181,346

8,149,136.3

9

43,554,9

82- (100,0

00)100,0

00 2,264

May-25

10.41

19,746,688

2,178,

000

17,568,68

8

100,952,03

0

4,633,42

9

164,213

71,487,391

10,615,322.

40

57,085,1

82- (100,0

00)100,0

00 2,264

Jun-25

10.50

19,850,621

2,178,

000

17,672,62

1

97,695,513

4,128,51

8

75,501

69,181,346

12,297,937.

61

66,524,8

82- (100,0

00)100,0

00 2,264

Jul-25

10.58

19,957,998

2,178,

000

17,779,99

8

100,952,03

0

33,526,1

01

8,958,999

71,487,391

10,862,765.

82

59,118,4

43

2,068,53

01,968,

530100,0

00 2,217

Aug-25

10.67

20,065,373

2,178,

000

17,887,37

3

100,952,03

0

40,233,7

59

13,105,293

71,487,391

9,617,300.6

4

52,656,3

34

20,630,0

5620,530,056

100,000 1,802

Sep-25

10.75

20,169,300

2,178,

000

17,991,30

0

97,695,513

19,898,3

97

4,378,596

69,181,346

8,536,797.7

4

47,011,9

66- (100,0

00)100,0

00 2,264

Oct-25

10.83

20,276,727

2,178,

000

18,098,72

7

100,952,03

0

16,976,6

91

2,840,771

71,487,391

6,441,776.8

4

35,686,5

51

7,253,77

37,153,

773100,0

00 2,101

Nov-25

10.92

20,380,740

2,178,

000

18,202,74

0

97,695,513

16,194,2

84

2,439,487

69,181,346

3,901,357.8

1

21,737,1

91

21,609,3

8921,509,389

100,000 1,764

Dec-25

11.00

20,488,292

2,178,

000

18,310,29

2

100,952,03

0

24,802,0

25

6,930,716

71,487,391

2,779,614.3

4

15,578,6

81

42,939,0

8542,839,085

100,000 1,302

Year

Month

Elapsed

Years

Impoundm

ent Tailin

gs Area, ft^2

Pond Area, ft^2

Beach Area,

ft^2


Monthly Balance at

TSF, gal


End of

Month

Water Storage, gal

Water Lost to

Tailings,

gpm

Tailings

Water, gal

Direct

Precip., gal

Runon f/

Undiv.Area, gal


Impoundment Free

Water Evap., gal

Impoundment

BeachEvap., gal

Document No.

Q400-05-028 25 July 2013

SGS FORM No. A263a-2/14/13 !


12

Jan-26

11.08

20,595,936

2,178,

000

18,417,93

6

100,952,03

0

22,735,6

51

5,605,986

71,487,391

2,680,636.9

7

15,112,2

74

40,113,3

65

40,013,365

100,000 1,365

Feb-26

11.16

20,693,257

2,178,

000

18,515,25

7

91,182,479

23,284,4

95

5,829,564

64,569,256

3,769,387.9

9

21,362,4

69

30,695,4

2430,595,424

100,000 1,503

Mar-26

11.25

20,801,128

2,178,

000

18,623,12

8

100,952,03

0

24,404,2

29

6,370,671

71,487,391

5,732,439.0

6

32,677,0

57

21,930,0

4221,830,042

100,000 1,772

Apr-26

11.33

20,905,658

2,178,

000

18,727,65

8

97,695,513

10,145,2

04

2,029,094

69,181,346

8,149,136.3

9

46,713,8

78- (100,0

00)100,0

00 2,264

May-26

11.41

21,013,836

2,178,

000

18,835,83

6

100,952,03

0

4,930,75

6

154,552

71,487,391

10,615,322.

40

61,202,4

70- (100,0

00)100,0

00 2,264

Jun-26

11.50

21,118,699

2,178,

000

18,940,69

9

97,695,513

4,392,25

2

71,034

69,181,346

12,297,937.

61

71,298,2

96- (100,0

00)100,0

00 2,264

Jul-26

11.58

21,227,257

2,178,

000

19,049,25

7

100,952,03

0

35,658,2

45

8,425,810

71,487,391

10,862,765.

82

63,338,7

26- (100,0

00)100,0

00 2,264

Aug-26

11.67

21,336,037

2,178,

000

19,158,03

7

100,952,03

0

42,781,6

10

12,320,526

71,487,391

9,617,300.6

4

56,396,8

77

18,652,5

9718,552,597

100,000 1,846

Sep-26

11.75

21,441,536

2,178,

000

19,263,53

6

97,695,513

21,153,5

44

4,114,782

69,181,346

8,536,797.7

4

50,336,3

66- (100,0

00)100,0

00 2,264

Oct-26

11.83

21,550,807

2,178,

000

19,372,80

7

100,952,03

0

18,043,4

14

2,668,488

71,487,391

6,441,776.8

4

38,198,7

44

5,636,02

05,536,

020100,0

00 2,137

Year

Month

Elapsed

Years

Impoundm

ent Tailin

gs Area, ft^2

Pond Area, ft^2

Beach Area,

ft^2


Monthly Balance at

TSF, gal


End of

Month

Water Storage, gal

Water Lost to

Tailings,

gpm

Tailings

Water, gal

Direct

Precip., gal

Runon f/

Undiv.Area, gal


Impoundment Free

Water Evap., gal

Impoundment

BeachEvap., gal

Document No.

Q400-05-028 25 July 2013

SGS FORM No. A263a-2/14/13 !


Nov-26

11.92

21,656,817

2,178,

000

19,478,81

7

97,695,513

17,208,2

38

2,290,571

69,181,346

3,901,357.8

1

23,261,0

45

20,950,5

7420,850,574

100,000 1,779

Dec-26

12.00

21,766,652

2,178,

000

19,588,65

2

100,952,03

0

26,349,5

39

6,504,688

71,487,391

2,779,614.3

4

16,666,3

29

42,972,9

2242,872,922

100,000 1,301

13

Jan-27

12.08

21,876,802

2,178,

000

19,698,80

2

100,952,03

0

24,149,5

87

5,258,917

71,487,391

2,680,636.9

7

16,163,2

50

40,129,2

56

40,029,256

100,000 1,365

Feb-27

12.16

21,976,580

2,178,

000

19,798,58

0

91,182,479

24,728,5

18

5,466,252

64,569,256

3,769,387.9

9

22,843,1

37

30,295,4

6730,195,467

100,000 1,513

Mar-27

12.25

22,087,382

2,178,

000

19,909,38

2

100,952,03

0

25,913,2

84

5,970,633

71,487,391

5,732,439.0

6

34,933,9

82

20,782,1

3420,682,134

100,000 1,798

Apr-27

12.33

22,194,963

2,178,

000

20,016,96

3

97,695,513

10,770,8

84

1,900,722

69,181,346

8,149,136.3

9

49,929,8

92- (100,0

00)100,0

00 2,264

May-27

12.41

22,306,511

2,178,

000

20,128,511

100,952,03

0

5,234,07

4

144,696

71,487,391

10,615,322.

40

65,402,7

05- (100,0

00)100,0

00 2,264

Jun-27

12.50

22,414,850

2,178,

000

20,236,85

0

97,695,513

4,661,82

5

66,469

69,181,346

12,297,937.

61

76,177,3

85- (100,0

00)100,0

00 2,264

Jul-27

12.58

22,527,218

2,178,

000

20,349,21

8

100,952,03

0

37,841,9

63

7,879,724

71,487,391

10,862,765.

82

67,661,0

94- (100,0

00)100,0

00 2,264

Aug-27

12.67

22,640,032

2,178,

000

20,462,03

2

100,952,03

0

45,396,2

95

11,515,174

71,487,391

9,617,300.6

4

60,235,5

41

16,623,2

6516,523,265

100,000 1,891

Year

Month

Elapsed

Years

Impoundm

ent Tailin

gs Area, ft^2

Pond Area, ft^2

Beach Area,

ft^2


Monthly Balance at

TSF, gal


End of

Month

Water Storage, gal

Water Lost to

Tailings,

gpm

Tailings

Water, gal

Direct

Precip., gal

Runon f/

Undiv.Area, gal


Impoundment Free

Water Evap., gal

Impoundment

BeachEvap., gal

Document No.

Q400-05-028 25 July 2013

SGS FORM No. A263a-2/14/13 !


Sep-27

12.75

22,749,648

2,178,

000

20,571,64

8

97,695,513

22,444,0

86

3,843,529

69,181,346

8,536,797.7

4

53,754,5

14- (100,0

00)100,0

00 2,264

Oct-27

12.83

22,863,392

2,178,

000

20,685,39

2

100,952,03

0

19,142,3

76

2,490,998

71,487,391

6,441,776.8

4

40,786,8

62

3,969,37

33,869,

373100,0

00 2,175

Nov-27

12.92

22,973,944

2,178,

000

20,795,94

4

97,695,513

18,254,8

11

2,136,864

69,181,346

3,901,357.8

1

24,833,9

20

20,270,5

6420,170,564

100,000 1,795

Dec-27

13.00

23,088,692

2,178,

000

20,910,69

2

100,952,03

0

27,949,9

29

6,064,104

71,487,391

2,779,614.3

4

17,791,1

41

43,007,9

1642,907,916

100,000 1,300

14

Jan-28

13.08

23,203,979

2,178,

000

21,025,97

9

100,952,03

0

25,614,6

44

4,899,300

71,487,391

2,680,636.9

7

17,252,2

24

40,145,7

22

40,045,722

100,000 1,364

Feb-28

13.16

23,308,587

2,178,

000

21,130,58

7

91,182,479

26,227,3

21

5,089,157

64,569,256

3,769,387.9

9

24,379,9

76

29,880,3

3629,780,336

100,000 1,523

Mar-28

13.25

23,424,951

2,178,

000

21,246,95

1

100,952,03

0

27,482,5

42

5,554,636

71,487,391

5,732,439.0

6

37,280,9

46

19,588,4

3119,488,431

100,000 1,825

Apr-28

13.33

23,538,126

2,178,

000

21,360,12

6

97,695,513

11,422,7

01

1,766,988

69,181,346

8,149,136.3

9

53,280,2

50- (100,0

00)100,0

00 2,264

May-28

13.41

23,655,675

2,178,

000

21,477,67

5

100,952,03

0

5,550,64

6

134,410

71,487,391

10,615,322.

40

69,786,4

85- (100,0

00)100,0

00 2,264

Jun-28

13.50

23,770,033

2,178,

000

21,592,03

3

97,695,513

4,943,67

5

61,695

69,181,346

12,297,937.

61

81,278,6

89- (100,0

00)100,0

00 2,264

Year

Month

Elapsed

Years

Impoundm

ent Tailin

gs Area, ft^2

Pond Area, ft^2

Beach Area,

ft^2


Monthly Balance at

TSF, gal


End of

Month

Water Storage, gal

Water Lost to

Tailings,

gpm

Tailings

Water, gal

Direct

Precip., gal

Runon f/

Undiv.Area, gal


Impoundment Free

Water Evap., gal

Impoundment

BeachEvap., gal

Document No.

Q400-05-028 25 July 2013

SGS FORM No. A263a-2/14/13 !


14Jul-28

13.58

23,888,842

2,178,

000

21,710,84

2

100,952,03

0

40,129,2

63

7,307,735

71,487,391

10,862,765.

82

72,188,4

89- (100,0

00)100,0

00 2,264

Aug-28

13.67

24,008,319

2,178,

000

21,830,31

9

100,952,03

0

48,139,8

94

10,670,114

71,487,391

9,617,300.6

4

64,263,4

66

14,493,8

8114,393,881

100,000 1,939

Sep-28

13.75

24,124,596

2,178,

000

21,946,59

6

97,695,513

23,800,5

67

3,558,416

69,181,346

8,536,797.7

4

57,347,3

07- (100,0

00)100,0

00 2,264

Oct-28

13.83

24,245,443

2,178,

000

22,067,44

3

100,952,03

0

20,299,4

98

2,304,114

71,487,391

6,441,776.8

4

43,511,9

51

2,214,52

32,114,

523100,0

00 2,214

Nov-28

13.92

24,363,081

2,178,

000

22,185,08

1

97,695,513

19,358,6

03

1,974,754

69,181,346

3,901,357.8

1

26,492,7

89

19,553,3

7719,453,377

100,000 1,811

Dec-28

14.00

24,485,372

2,178,

000

22,307,37

2

100,952,03

0

29,640,6

75

5,598,645

71,487,391

2,779,614.3

4

18,979,4

59

43,044,8

8642,944,886

100,000 1,299

Jan-29

14.08

24,608,424

2,178,

000

22,430,42

4

100,952,03

0

27,164,9

98

4,518,745

71,487,391

2,680,636.9

7

18,404,5

99

40,163,1

46

40,063,146

100,000 1,364

Feb-29

14.16

24,720,239

2,178,

000

22,542,23

9

91,182,479

27,815,7

42

4,689,514

64,569,256

3,769,387.9

9

26,008,7

07

29,440,3

8429,340,384

100,000 1,534

Mar-29

14.25

24,844,793

2,178,

000

22,666,79

3

100,952,03

0

29,148,3

24

5,113,051

71,487,391

5,732,439.0

6

39,772,2

72

18,321,3

0318,221,303

100,000 1,853

Apr-29

14.33

24,966,107

2,178,

000

22,788,10

7

97,695,513

12,115,6

79

1,624,809

69,181,346

8,149,136.3

9

56,842,1

78- (100,0

00)100,0

00 2,264

Year

Month

Elapsed

Years

Impoundm

ent Tailin

gs Area, ft^2

Pond Area, ft^2

Beach Area,

ft^2


Monthly Balance at

TSF, gal


End of

Month

Water Storage, gal

Water Lost to

Tailings,

gpm

Tailings

Water, gal

Direct

Precip., gal

Runon f/

Undiv.Area, gal


Impoundment Free

Water Evap., gal

Impoundment

BeachEvap., gal

Document No.

Q400-05-028 25 July 2013

SGS FORM No. A263a-2/14/13 !


15

May-29

14.41

25,092,287

2,178,

000

22,914,28

7

100,952,03

0

5,887,73

8

123,457

71,487,391

10,615,322.

40

74,454,4

06- (100,0

00)100,0

00 2,264

Jun-29

14.50

25,215,210

2,178,

000

23,037,21

0

97,695,513

5,244,24

2

56,605

69,181,346

12,297,937.

61

86,718,7

53- (100,0

00)100,0

00 2,264

Jul-29

14.58

25,343,088

2,178,

000

23,165,08

8

100,952,03

0

42,572,1

53

6,696,837

71,487,391

10,862,765.

82

77,023,8

53- (100,0

00)100,0

00 2,264

Aug-29

14.67

25,471,858

2,178,

000

23,293,85

8

100,952,03

0

51,074,4

85

9,766,227

71,487,391

9,617,300.6

4

68,571,7

89

12,216,2

6112,116,261

100,000 1,990

Sep-29

14.75

25,597,341

2,178,

000

23,419,34

1

97,695,513

25,253,5

31

3,253,024

69,181,346

8,536,797.7

4

61,195,6

46- (100,0

00)100,0

00 2,264

Oct-29

14.83

25,727,920

2,178,

000

23,549,92

0

100,952,03

0

21,540,7

02

2,103,651

71,487,391

6,441,776.8

4

46,435,0

57

332,158

232,158

100,000 2,256

Nov-29

14.92

25,855,190

2,178,

000

23,677,19

0

97,695,513

20,544,2

14

1,800,628

69,181,346

3,901,357.8

1

28,274,6

21

18,783,0

2918,683,029

100,000 1,829

Dec-29

15.00

25,987,652

2,178,

000

23,809,65

2

100,952,03

0

31,459,2

54

5,097,993

71,487,391

2,779,614.3

4

20,257,6

22

43,084,6

5142,984,651

100,000 1,299

Average

99,052,395

14,912,0

366,122

,75970,142,198

6,220,115

28,710,7

17

20,305,5

1320,205,513

100,000 1,800

Year

Month

Elapsed

Years

Impoundm

ent Tailin

gs Area, ft^2

Pond Area, ft^2

Beach Area,

ft^2


Monthly Balance at

TSF, gal


End of

Month

Water Storage, gal

Water Lost to

Tailings,

gpm

Tailings

Water, gal

Direct

Precip., gal

Runon f/

Undiv.Area, gal


Impoundment Free

Water Evap., gal

Impoundment

BeachEvap., gal

Document No.

Q400-05-028 25 July 2013

SGS FORM No. A263a-2/14/13 !


Water inflows include the process water contained in the tailings slurry, direct precipitation on the tailings surface, and surface water runon from external areas that cannot be diverted.

The solids content of thickened backfill plant rejects is assumed to be 50 percent. The solids content of the backfill plant by-pass whole tailings is assumed to be 60 percent. An average of 2,261 gpm of water will be delivered to the tailings impoundment with the tailings slurry. The water balance analysis for the TSF indicates an average of approximately 462 gpm will be recovered from the TSF. Reclaim rates will vary subject to seasonal and climatic factors. An average of approximately 1,800 gpm of makeup water will be required to offset TSF losses.

Additional water will be entrained with the mine backfill. The cyclone overflow used for mine backfill will have a solids content of approximately 70 percent by weight, and will include 223 tons per hour (tph) of water. Approximately 50 percent of the water delivered with the backfill to the underground will be entrained with the backfill. Backfill drainage will be collected and returned to the surface. 18.2 summarizes estimate water usage rates for the proposed 25,000 ton per day mining operation.

18.5 Tailings Impoundment

18.5.1 Location

The PEA level layout of the proposed tailings storage facility (TSF) is shown on Figure 18.1. The facility is located on state land in Sections 5, 6, 7, and 8 in T8S, R18E, approximately 4.2 miles west of the Saloon Gulch shaft site, and is just north of the existing Copper Creek Road alignment. The elevation at the TSF site ranges from 3590

Table 18.2 Summary of Water Uses, Copper Flat Project, 25,000 tpd Mining Rate

Use DescriptionQuantity,

gpm

TSF Entrained water and evaporative losses 1800

Underground Mine Water entrained in hydraulic backfill (50% of water delivered) 330

Process Plant Misc losses, dust control, gland seal water, shop usage 25

Potable Water 15

Sanitary Facilities Sewage disposal, miner’s dry 25

Underground Mine Dust suppression, ramp maintenance 675

Total Total make-up from raw water supply 2,870

Total Total gallons per ton mined at 25,000 tpd 165


SGS FORM No. A263a-2/14/13 !


to 3950 feet above mean sea level (amsl). This location allows the utilization of four contributing “fingers” within Well Canyon. A relatively small saddle dam will also be required for later phases (3 through 5) on the northernmost finger.


SGS FORM No. A263a-2/14/13 !


! Figure 18.1 - Tailings Storage Facility


SGS FORM No. A263a-2/14/13 !


18.5.2 Tailings Characteristics

The TSF will be used to contain tailings whose characteristics change subject to backfill plant operation. While mine backfill is in progress, the flotation tailings stream will be passed through cyclones. The coarse sand fraction (cyclone underflow) will be amended with Portland cement and used for underground mine hydraulic backfill. The cyclone overflow (backfill plant rejects) will be thickened to an anticipated solids content of 50 percent and deposited in the TSF. In order to fill the underground voids with hydraulic fill, backfill plant utilization will need to be on the order of 74 percent. The backfill plant rejects will be delivered to the TSF at an average rate of 9,250 tpd.

While the backfill plant is not in operation, the flotation tailings stream will by-pass the cyclones and be routed directly to the thickener and TSF. Based on backfill plant utilization requirements, this will occur approximately 26 percent of the time. It is assumed that the backfill plant by pass will be delivered to the TSF at a solids content of approximately 60 percent by weight at an average rate of 6,500 tpd. The TSF will receive approximately 15,750 tons of tailings solids per day.

18.5.3 TSF Description

The proposed TSF will consist of a Phase 1 starter dam constructed to the elevation of 3,760 feet amsl and four downstream dam raises (phases 2 through 5) constructed with locally borrowed fill material. Table 18.3 illustrates elevation versus capacity and rate of rise relationships for the proposed TSF. The site is located on alluvial valley fill and is not anticipated to require drilling and blasting to generate local borrow material. It is assumed that local borrow materials will be used to construct the dam; however, development waste from the underground mine and overburden generated from the Old Reliable satellite pit could potentially be used as fill. Use of mine waste for dam fill could potentially reduce capital construction costs.


SGS FORM No. A263a-2/14/13 !


!


SGS FORM No. A263a-2/14/13 !


Common TSF embankment construction practice at copper mining operations in the southwest involves earthen starter dam construction with subsequent raises constructed by the upstream method using tailings sand. Adoption of the upstream raise method is contingent on the development of a well-drained embankment and upper tailings beach to support raise foundation loads. The downstream raise method is considered for the Copper Creek mining operation because much of the material discharged into the TSF will consist of rejects (cyclone overflow slimes) from the backfill plant which are anticipated to exhibit higher moisture content, lower permeability and slower rates of consolidation relative to the whole tailings disposed of at most base metal mining projects where hydraulic backfill is not employed. As such, foundation conditions suitable for the construction of upstream embankment raises cannot be assumed with confidence. Alternative embankment raise options, which could potentially reduce earthworks costs if conditions warrant their use, are discussed in Section 26. Note that the rate of tailings rise will be under 10 ft per year for the proposed TSF after Year 4. Low rates of rise may facilitate alternative raise construction methods.

Tailings will be discharged into the TSF via multiple spigot points located on the crest of the tailings embankment and around the TSF periphery. The TSF will include a system of underdrains (finger drains) constructed in natural drainage channels on the TSF floor. Underdrains are assumed to consist of perforated, corrugated polyethylene pipe (PCPE) placed in a drain-fill envelope. Underdrainage will be routed to a seepage collection pond located downstream of the TSF embankment.

18.5.4 Surface Water Management

Due to the nature of the topography within Well Canyon, it is generally impractical to divert external area runoff away from the TSF. The site does have a limited upstream contributing area; however, it is anticipated that sufficient freeboard will be maintained in the TSF to accommodate storm water associated with the design storm event. The TSF site is remote from infrastructure and human habitations. Due to the potential for environmental damage, a design storm based on probable maximum precipitation is assumed. At closure, a spillway will be required that is capable of passing the peak discharge associated with the design storm event.

18.5.5 Water Reclaim

The water reclaim system will consist of barge mounted pumps inside the TSF and barge mounted or submersible turbine pumps located in the seepage reclaim pond. It is assumed that both systems will be capable of transporting water back to the processing facilities at a rate that is greater than or equal to the average TSF process water inflow rate. For this PEA, the process water inflow rate is estimated to be an average of approximately 2,300 gpm.


SGS FORM No. A263a-2/14/13 !


18.6 Diversion Tunnel

Underground mining is predicted to result in subsidence of the ground surface over the underground stopes. Cracks could propagate to the ground surface and could impact the flow of surface water in Copper Creek. The PEA considers the construction of a diversion tunnel to intercept surface flow in Copper Creek upstream of the subsidence area. Surface flows will be released in Dark Canyon below the mine disturbance area and the subsidence crack limit, and above the confluence of Dark Canyon and Copper Creek.

The diversion tunnel preliminary design storm is the 100-year, 24-hour event. Components of the Copper Creek diversion include an earthen check dam in Copper Creek upstream of the projected crack limit. The purpose of the check dam is to attenuate flood flows associated with major storm events, and generate head at the tunnel inlet. The inlet to the diversion is assumed to be a concrete headwall/wingwall structure with a trash rack/security grate. Based on a PEA level tunnel layout, the tunnel length is estimated to be 7,540 feet. Tunnel diameter will be twelve feet.

Tunneling is assumed to be completed with a tunnel boring machine. It is anticipated that the majority of the tunnel route will be in competent granodiorite; however, diversion tunnel costs include structural lining.

A concrete outlet structure and security grate will be constructed at the tunnel outlet in Dark Canyon.

18.7 Hydraulic Backfill Plant Description

The Hydraulic Backfill Plant considered in this study is designed based on a mining rate of 25,000 tons/d. The proposed Hydraulic Backfill Plant will produce sufficient hydraulic backfill to fill 85 percent of the void in the underground mine. It is assumed the remainder of the void will be filled by waste rock or it is not required for filling at all. A process flow diagram (PFD) which illustrates the hydraulic backfill process can be found in the following Figure 18.2.


SGS FORM No. A263a-2/14/13 !


!

Figure 18.2 - Hydraulic Backfill Plant Process Flowsheet


SGS FORM No. A263a-2/14/13 !


The Hydraulic Backfill Plant will process all mill tailings sent from the mill tailings pump-box. Hydraulic backfill is prepared by processing mill tailings through hydrocyclones to remove the fine tailings solids (referred to as slimes). Approximately 50 wt% of solids in the tailings are removed as slimes. The slimes from the hydrocyclones overflow will be thickened and transported to the Tailings Storage Facility (TSF). The hydrocyclones also act as a dewatering process to produce dewatered coarse tailings in the underflow. The coarse particles from the cyclone underflow (referred to as sands) are mixed with binder in an agitated mix tank to a hydraulic backfill consistency. The resulting hydraulic backfill is discharged to the underground stopes via a underground piping distribution system.

The Hydraulic Backfill Plant will be capable to produce hydraulic backfill with different UCS by adjusting the amount of cement addition to the backfill, to accommodate the strength requirements in the stopes. For this study, it is assumed an average of 5 percent binder content is required for backfill.

The Hydraulic Backfill Plant contains various equipment and associated instrumentation that will operate in sequence to produce the hydraulic backfill for use in the underground. The major components include the hydrocyclones, binder system, mixing tank and the discharge pumps to the underground distribution system. It is assumed the Hydraulic Backfill Plant will be located in close proximity to the mill. It is anticipated that compressed air required for the backfill plant will be supplied by the mill’s compressed air system.

Hydrocyclones

A cyclopack consisting of multiple hydrocyclones will be utilized to remove the fine particles and water in the tailings through the overflow. The slurry will enter the hydrocyclones at approximately 25 wt% solids. Finer particles and the majority of the water will leave as overflow. The thickened coarse tailings will exit from the underflow at approximately 70 wt% solids.

Binder System

The binder system will add binder at a rate proportional to the weight of solids in the slurry. The rate can also be varied in order to meet the strength requirement of the backfill.

Binder will be delivered to site by trucks and transfer to the silos pneumatically. The binder system will consist of two large binder storage silos, with a total capacity to hold 2 days of cement usage. Normal Portland Cement (NPC) has been considered as the binder in this study. Alternatively, a blend of NPC, slag cement or fly ash can be used to reduce binder operating costs.


SGS FORM No. A263a-2/14/13 !


Mixing Tank

Thickened coarse tailings from the hydrocyclones underflow, binder, and water will be continuously added into the mixing tank during operation. The mixing tank will blend the materials to a consistency suitable for hydraulic backfill. The hydraulic fill pumps will discharge the hydraulic backfill to the underground stopes. It is expected the mixing tank will have a retention time of one minute.


SGS FORM No. A263a-2/14/13 !


19.0 MARKET STUDIES AND CONTRACTS

Redhawk is not currently in production and has no sales contracts in place at this time. Should the project go into production, smelter agreements for treatment and refining of copper and molybdenum concentrates will be put into place. For the purpose of this PEA study, smelter terms and costs have been estimated.


SGS FORM No. A263a-2/14/13 !


20.0 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT

20.1 Hydrogeological

Water Resources Review and Water Supply

Past investigation data compiled by Water Management Consultants (WMC) of Tucson, Arizona were reviewed for the purposes of evaluating the potential for a continuous supply of groundwater at a rate of from 400 to 500 gpm from a location near the project area. The Water Management report describes the groundwater yield potentials for the various rock types located throughout the region and concluded that the basin fill sediments assigned to the Gila Conglomerate would provide the greatest groundwater production potential. As a result of this interpretation four test wells were installed by WMC at various locations in the Gila Conglomerate. These test well locations are west of the project area by less than 1/2 mile.

A 48 hour aquifer test was conducted at Test Well WW-4R and Test Well WW-1R was used as an observation well. Data from these tests suggest that the aquifer is capable of producing sufficient short-term discharge rates for the planned operation. The specific capacity, discharge (gpm) divided by drawdown (ft), value of 5.1 calculated from the test data suggests that the aquifer is capable of providing approximately 5 gpm for every ft of drawdown during short term pumping. The significant caveats to this interpretation are that the aquifer test was performed for only 48 hours, at a discharge rate of only 50 gpm, and that the test and observation wells are located fairly close to the range-front fault system. Range-front fault systems generally behave as aquitards and can significantly impede groundwater flow to nearby wells.

In light of this preliminary and limited testing Golder Associates was authorized to conduct a regional hydrogeological investigation and well siting study. The complete technical memorandum is provided in Appendix 7-A. The hydrogeological study was expanded from the area studied by WMC to include other major hydrogeological regions of the San Pedro river basin including areas with historic high yield and deep production wells.

As is typical for deep basins throughout Arizona, groundwater movement is from higher elevations in the mountains towards the valley floor. In the San Pedro Valley the groundwater flows towards the center of the basin and then northwest parallel to the San Pedro River, as indicated by the Arizona Department of Water Resources (ADWR) well and water depth information. The primary basin-fill aquifer system is separated from the fractured bedrock aquifer by a west-dipping, mountain bounding fault system. The mountain bounding fault is thought to act as a limited recharge feature in select areas but generally acts as an aquitard impeding the movement of groundwater. Groundwater moves readily between the younger and older basin-fill units, and between the streambed alluvium and the basin fill units.


SGS FORM No. A263a-2/14/13 !


Gradients calculated from groundwater elevation data indicate that gradients within the central unconsolidated river gravels are approximately 0.004 ft per ft (ft/ft). Steeper gradients of 0.06 and 0.03 ft/ft are present within the basin-fill and alluvial fan sediments, respectively. Areas with lower groundwater gradients have relatively wider spaced contour lines than areas with steeper gradients. Areas with relatively flatter gradients are indicative of regions with higher hydraulic conductivity values, leading to potentially higher groundwater production values from correctly designed and constructed production wells.

Mountain front recharge is thought to be the main source of recharge for the basin fill aquifer systems. Recharge to the basin-fill aquifer systems is also thought to occur as infiltration from streambed alluvium. Direct infiltration on mountain front pediments is considered negligible due to high evaporation rates and low rainfall totals.

The lower San Pedro river basin contains two major water-bearing units: the streambed Holocene alluvium that forms the San Pedro River channel and floodplain, and the alluvial basin fill deposits that are composed of a younger basin-fill, older basin-fill, and a basal conglomerate. The basin-fill deposits form the basin’s primary aquifer because of their high permeability and large groundwater storage. However due to recent surface water adjudication, water rights, and permitting issues, groundwater from the upper San Pedro River aquifer including its base flow (Holocene alluvium) should not be considered as potential supply water for the Copper Creek project. The Golder study concentrated on the lower basin fill groundwater that is believed to be hydrologically isolated from the San Pedro River Holocene alluvial system.

Regionally, the hydrologic characteristics of the lower basin fill aquifer can vary widely due to compaction and the presence of fine-grained units within the basin-fill sediments. Due to this variability of basin-fill sediments, the groundwater yields from wells drilled in the basin fill along the east side of the valley range from yields of 70 gpm to as much as 1,900 gpm. Well records indicate that the wells with high yields are pumping from both the lower and upper aquifers. Towards the center of the valley, groundwater from wells within the lower aquifer system are found to exhibit artesian conditions. Most wells drilled deeper than 500 ft encounter artesian conditions especially in sand and gravel layers 600 to 800 and 1,200 to 1,300 ft below ground surface (bgs). Away from the river valley towards the Galiuro Mountains information on groundwater is poor.

As a result of the hydrogeological well sitting study four locations had been identified with hydrogeological and logistical characteristics thought to be suitable for the siting of production wells. However, due to adjudication of groundwater within the basin, ADWR is reluctant to provide groundwater exploration permits within a three-mile radius of the active channel of the San Pedro River.

The existence of two previously installed wells along Copper Creek Road provided an opportunity to obtain hydrologic test data without drilling new groundwater exploration holes. The two wells (55-615551 and 55-615552) were drilled and


SGS FORM No. A263a-2/14/13 !


constructed approximately 450 ft apart in Sections 14 and 15 of Township 8 South, Range 17 East. The boreholes were drilled to depths of 360 and 2,074 ft below ground surface respectively (See Appendix 7-A Figure 3). The land is currently leased to the Mercer Ranch and a stock water pump is installed in well 55-615552. Discussions with local residents indicate that this well provided water to past operations in Copper Creek and was pumped at 600 gallons per minute (gpm). Golder conducted a constant-rate aquifer test at well 55-615552 on December 8, 2011. The well was pumped for 24 hours at an average rate of 294 gpm and recorded a total drawdown of 162 ft bgs. Analysis of the test data were used to estimate aquifer parameters of the basin fill sediments within the immediate area of the well 55-651552. The aquifer test technical memorandum is provided as Appendix 7-B.

Using the estimated aquifer parameters of the basin fill sediments from the well test, Golder conducted a well spacing study in the area of the existing Right-Of-Way (ROW) for the Copper Creek project. Analytical modeling was performed to simulate the drawdown from multiple pumping wells to evaluate the overall spacing needed between future wells, assuming a planned aquifer withdrawal rate of 800 gpm per well, or 5,600 gpm total. The drawdown analysis was performed using AQTESOLV Pro © (v.4.5) (Duffield, 2007), which has the capability of simulating drawdown from multiple pumping wells using the Theis equation and assumes a flat, infinite aquifer. The aquifer test technical memorandum is provided as Appendix 7-C. From the studies discussed above and field exploration, a total of five sites have been selected for further investigation of groundwater production. A sixth site, for a piezometer only, has been proposed for stratigraphic investigation. The purpose is to evaluate the proposed hydraulic disconnection between the deep aquifer and the Holocene alluvial aquifer of the San Pedro River system. This site is between the three-mile buffer zone and the active channel of the San Pedro River.

The five sites selected for further investigation of groundwater production were sited based on several factors. First, the sites had to be outside of the three-mile buffer zone of the active channel of the San Pedro River. Second, the sites had to be as near the existing ROW as possible for access to future water pipeline routes. Lastly, the sites should be as near to the break in pediment elevation (indicating a possible buried fault) as possible.

All six well sites have been submitted to the Arizona State Land Department (ASLD) for inclusion into a Supplemental Land Use Permit (SLUP). The SLUP is an extension to the ROW which allows Redhawk to conduct groundwater exploration drilling at the proposed sites.

Table 20.1 depicts the estimated costs associated with the recommended groundwater test/production wells and piezometer well installation programs.

Table 20.1 Hydrogeological Cost Estimates


SGS FORM No. A263a-2/14/13 !


20.2 Geochemical Considerations

20.2.1 Properties of Development Rock

An analysis of the geochemical properties of the mine development rock from the Mammoth and Childs-Aldwinkle breccia areas has been performed. Test results indicate that the country rock in the vicinity of the initial breccia mining operation has a low potential for acid generation, and drainage from the development rock is expected to be of good quality. Similar testing has not been performed for development rock in the vicinity of the deeper porphyry mineralization of the Keel and American Eagle deposits. Future development rock is likely to be consumed in mining backfill operations to increase backfill strength and to develop a surface layer capable of supporting mining equipment traffic for room and pillar mining. At present, a permanent surface storage facility for development rock is not being considered; however, geochemical characterization of development rock in accordance with Arizona Mining Guidance standards is likely to be required as a condition of permitting.

20.2.2 Geochemical Properties of Tailings

To date, an analysis of the geochemical properties of the future tailings has not been performed. The material is expected to contain residual pyrite and could exhibit acid generating potential. While the arsenic mineral tennantite is also known to occur at Copper Creek, its distribution in the initial breccia mining targets is estimated to be limited. While tennantite is reported to be insoluble under neutral to alkaline pH conditions, oxidation of tennantite following tailings disposal could result in the

Work Task DescriptionApproximate Cost

US$

Aquifer Test/Production Wells - 2

Oversight of drilling, construction, and testing of two San Pedro Valley lower aquifer test wells. Installed within the lower basin fill sequence. Would include zonal water quality testing, aquifer testing, and completion as 10-inch aquifer test/production well.

$125,000

Estimated Drilling Subcontractor Costs ≈$800,000 (2 wells)

Piezometer Wells - 3 Permitt ing and oversight of dri l l ing, construction, and testing of three standpipe piezometer wells. Two piezometers would be paired in the vicinity of the test wells above for aquifer data acquisition. One piezometer to be installed between three-mile buffer and active channel to show aquifer disconnection from alluvial aquifer.

$70,000

Estimated Drilling Subcontractor Costs ≈$300,000 (2 wells)

Misc. hydrogeological investigations

Investigation of status of state land well and design of BHP wells

$12,000


SGS FORM No. A263a-2/14/13 !


mobilization of arsenic. The potential for acid generation and/or leachable arsenic could lead to a requirement for lined tailings disposal facilities to mitigate impacts to water resources.

20.2.3 Geochemical Characterization Costs

Initial, pre-start-up characterization of process tailings, development rock and process water will be required in accordance with Arizona Mining Guidance Tier 1 and 2 testing standards. The need to conduct extended kinetic testing to evaluate acid generating potential is assumed. A geochemical characterization study can be expected to require 30 to 40 weeks to complete if conventional kinetic testing is undertaken. The time required to complete geochemical characterization studies could be considerably longer if the bench and pilot scale studies must be conducted to produce representative samples.

20.3 Environmental Permitting

The discussion below provides a comprehensive discussion of potential environmental permitting considerations for the project. Some key considerations with regard to cost and schedule include:

▪ Proximity (less than 2 miles) of the Galiuro Wilderness Area, a Class I air shed

▪ Significant historic mining features throughout the project area

▪ Significant NEPA review requirements and associated public involvement

In addition, the project has several favorable attributes from a permitting perspective, including proximity to existing mining districts and associated infrastructure, and remoteness from residential and urban centers.


SGS FORM No. A263a-2/14/13 !


BLM Approvals

Because significant portions of the project are located on lands managed by the Bureau of Land Management (BLM), it is assumed for this evaluation that a Mine Plan of Operations (MPO) will be required for development of the project. Once the MPO has been submitted, the BLM will begin its review process under NEPA.

NEPA analysis includes a number of resource areas (endangered species, historic properties, air, water, etc.) and requires public involvement. In general, a given project will fall into one of three categories, described below in ascending order of difficulty level.

▪ Categorical exclusion - Projects in this category are small or insignificant enough that an extensive public review process is not required. For categorical exclusions, the BLM is required to document that the impacts of the proposed project will not result in significant effects to the environment, and that the activity conforms with a prescribed set (“category”) of low-impact actions.

▪ Environmental Assessment (EA) - An EA requires a full analysis of impacts to a host of environmental resource areas, as well as public scoping to determine additional potential environmental resources to be evaluated. In order for a project to be approved, the EA must result in a Finding of No Significant Impact (FONSI) to the evaluated environmental resources. Note that a FONSI can be achieved through mitigation efforts proposed by the project proponent.

▪ Environmental Impact Statement (EIS) - An EIS is an expanded EA, for projects in which a FONSI is not anticipated to be acquired. The EIS results in a Record of Decision (ROD) outlining the evaluated impacts of the project, along with the mitigation responsibilities of the project proponent.

Given the size of the project, it is assumed for the purposes of this analysis that an EIS will be required. Following submittal of the MPO, public scoping is initiated by the lead federal agency (in this case, BLM) in order to identify potential environmental concerns. The impacts of the project are then evaluated for a suite of environmental resources, both from a standard BLM list and those resources identified during scoping. These resource areas include, but are not limited to:

▪ Access and recreation ▪ Agriculture and grazing ▪ Visual resources ▪ Air quality ▪ Geology ▪ Soils ▪ Groundwater ▪ Surface water


SGS FORM No. A263a-2/14/13 !


▪ Vegetation ▪ Wildlife ▪ Special interest species ▪ Cultural resources ▪ Socioeconomic resources ▪ Infrastructure ▪ Traffic/transportation

Alternatives to the proposed project are developed as part of the EIS process, and are often derived as a result of comments received during scoping. These alternatives are all evaluated for the same environmental parameters. In addition, because the approval of a Mine Plan of Operations is a federal action, compliance with the Endangered Species Act (ESA, Section 7) and National Historic Preservation Act (NHPA, Section 106) is required (see below).

Given the potentially significant processing times for the NEPA effort, it is advisable to approach the BLM as early in the permitting process as possible.

Clean Air Act

Air quality is regulated at the federal level by the EPA under the Clean Air Act (CAA). The CAA requires maintenance of ambient air quality, as indicated by concentrations of six “criteria pollutants” (ozone, carbon monoxide, nitrogen dioxide, sulfur dioxide, particulate matter less than 2.5 microns and less than 10 microns aerodynamic diameter [PM2.5 and PM10], and lead). National Ambient Air Quality Standards (NAAQS) have been established for each of the criteria pollutants. The CAA requires that each state develop a State Implementation Plan (SIP) describing how these standards will be met. Air quality control regions (AQCRs) within each state are designated to assist in evaluating ambient air quality. Each AQCR is evaluated to determine if the NAAQS are met, and areas not meeting the standards are designated as “non-attainment” for the particular criteria pollutant that exceeds the relevant standard. The SIP must explain how the non-attainment areas will be managed in order to meet any standards not attained.

ADEQ has also developed a SIP to address regional haze, in concert with other states in the southwestern U.S. The regional haze SIP was finalized and submitted to the EPA on March 2, 2011, and includes the Galiuro Wilderness Area (a Class I airshed).

Air emissions are regulated under the CAA in the context of the NAAQS. The law and regulations differentiate between mobile and stationary sources, as well as between new and existing facilities. Mobile sources are typically vehicles, and regulated emissions from vehicles include both engine (exhaust) emissions and dust generation (from, for example, haul truck operation on dirt roads). Exhaust emission standards have been established by the EPA based on the NAAQS (with more stringent standards adopted by some states, but not Arizona). New or modified existing stationary sources must meet performance standards, referred to as New Source Performance Standards


SGS FORM No. A263a-2/14/13 !


(NSPS), established by the EPA for certain categories of sources. The standard of performance for a particular facility is based on the application of the best available system of emission reduction, taking into consideration cost. New major sources are subject to preconstruction review, with different standards and levels of review applied to facilities proposed within attainment areas (“Prevention of Significant Deterioration” requirements) and non-attainment or non-classifiable areas (“New Source Review” requirements).

The permitting components of the CAA for stationary sources are described in Title V of the Act; thus, air emission operating permits are commonly referred to as Title V permits. These permits comprehensively address all relevant air emissions limitations, and monitoring and reporting requirements. Mining operations qualify as Class I major sources, and must be permitted accordingly.

The project area is located within or very near proximity to the San Manuel Sulfur Dioxide Non-Attainment Area, designated as a result of emissions from BHP’s copper smelter stack and fugitive emissions. Operations at the San Manuel smelter ceased in 2002. In June 2002, ADEQ submitted to EPA the San Manuel Sulfur Dioxide Nonattainment Area State Implementation and Maintenance Plan showing attainment of the federal standard was reached and requesting redesignation to attainment. Included in the plan are emission limits for the smelters as codified in Arizona Administrative Code R18-2-715 and 715.01. In 2007, this area was redesignated to attainment.

The project area is less than two miles from the Galiuro Wilderness Area, a Class I airshed. The Title V permit application for this facility would be subject to preconstruction review under the New Source Review provisions, and the permit would have to comply with the Arizona regional haze SIP.

For the Copper Creek project, it is anticipated that CAA permitting would consist of application for a new Title V permit for a new mill; no other aspects of the project are anticipated to require permitting of air emissions. Modeling emissions will require site meteorological data (including wind rose data), along with mill throughput and emission control data. The modeling exercise can only be accomplished through utilization of technical data from a specified project plan; modeling based on surrogates can be useful for identifying key issues but does not substitute for site-specific data and equipment specifications. Thus, a highly developed project plan is required to complete this analysis.


SGS FORM No. A263a-2/14/13 !


CWA Section 404 Permit A CWA Section 404 permit is required if a project will result in the discharge of dredged or fill material into jurisdictional waters of the U.S. (waters). The U.S. Army Corps of Engineers (Corps) administers the Section 404 program. The definition of waters is broad; the Corps has commonly asserted jurisdiction over the typically dry arroyos or washes found throughout Arizona. Guidance related to the identification of waters was promulgated in 2007 (and revised in 2008) which has had the effect of limiting the scope of Corps jurisdiction over some ephemeral washes. A jurisdictional waters determination will be completed for the Copper Creek Project area to determine the extent of Corps jurisdiction, though it is anticipated that at the very least Copper Creek itself is likely jurisdictional.

Two primary Section 404 permitting avenues are available to project proponents: the Individual Permit (IP) and the Nationwide Permit (NWP). NWPs are available for certain specified categories and sizes of disturbance that result in only “minimal impact to the aquatic environment”. Individual Permits are required for larger projects or projects whose activities are not covered by the NWP program. The effort to obtain a Section 404 permit varies considerably both in time and cost, depending on the type and extent of the impacts. An Individual Permit can take up to two years and cost between US$ 100,000 and US$ 300,000. A NWP, on the other hand, can be accomplished in six to nine months, generally for less than US$ 20,000. The need for either permit type constitutes a federal nexus that requires compliance with the ESA and NHPA (see below).

For the purposes of this evaluation, given the proposed impacts to Copper Creek, it is reasonable to assume that an Individual Permit will be required for the project. As described above, the Individual Permit process is, in general, more time- and labor-intensive than the process for NWPs. The following steps are required to obtain the Individual Permit.

▪ CWA Section 404 Application - The Corps requires completion and submittal of Form ENG 4345, the CWA 404 permit application form. Supplementary information, including a complete project description and graphical representation of proposed activities and their relation to waters, is usually required as part of the application package.

▪ 404(b)(1) Alternatives Analysis - A 404(b)(1) Alternatives Analysis is an evaluation of the proposed project’s purpose and need, and demonstration that it is the least environmentally damaging practicable alternative (LEDPA). Development of alternatives for the 404(b)(1) analysis would inform the development of the BLM's EIS (see above).

▪ Habitat Mitigation and Monitoring Plan (HMMP) - The Corps requires the applicant to provide a plan describing compensatory mitigation for the loss of waters. Generally, the project proponent may mitigate for loss of waters by replacing them with constructed wetland projects or planned enhancements


SGS FORM No. A263a-2/14/13 !


of riparian areas, or by negotiating a fee in-lieu of mitigation arrangement. Current rule requires the Corps and the applicant to first consider a mitigation bank (though none currently exists in Arizona) or an in-lieu fee payment (in-lieu fee projects in Arizona are currently being developed and approved).

▪ Public Notice - For Individual Permits, the Corps is required to publish a Public Notice (PN) describing the project and its anticipated impacts to environmental resources. Public comments are solicited during a 15- to 30-day public comment period which follows the publication of the PN. The applicant is then required to draft a response to any comments received by the Corps as part of the PN process.

▪ NEPA Compliance - Preparation of an Environmental Assessment (EA) (or, less commonly, an EIS) is required by the Corps to satisfy their obligation for environmental review under NEPA. Review and incorporation of responses to comments received during the required public review process is required. This analysis assumes that the Corps’ obligations will be achieved through a cooperating agency arrangement with the BLM’s NEPA process.

▪ CWA Section 401 State Water Quality Certification - Under Section 401 of the CWA, the Arizona Department of Environmental Quality (ADEQ) reviews proposed projects for water quality compliance. ADEQ may grant or deny certification for any Section 404 permit based on the anticipated water quality impact of the project, and typically provides conditions for protecting water quality that the Corps then incorporates into the 404 permit.

Once the jurisdictional waters determination has been completed and approved by the Corps, the scope of the Section 404 permitting effort will be determined. Again, it is anticipated that an individual permit will be required.

Endangered Species Act Compliance

Section 7 of the ESA requires that, for any federally permitted action, the permitting authority must evaluate the potential impact of a project to federally listed species. If the lead federal agency determines that the project may affect a listed species, consultation with the United States Fish and Wildlife Service (USFWS) will be required. For the development of the Copper Creek Project, consultation with USFWS would be required for either the CWA Section 404 permit or the Mine Plan of Operations for the BLM. It is anticipated that the BLM would act as the lead federal agency for the purpose of the Section 7 consultation.

Even in the absence of a federal nexus, listed species are protected under the Section 9 ‘take’ provision, which prohibits the harming, harassing, injuring, or killing of a federally listed threatened or endangered species. However, take applies only to wildlife (property of the state), not plants (which are owned by the property owner). In the absence of a federal nexus, federally listed plants on private property are protected by state and local plant protection laws and ordinances.


SGS FORM No. A263a-2/14/13 !


Given the location of the project area and the nature of the onsite habitat, three federally listed species are anticipated to have a reasonable potential to occur on the project area: southwestern willow flycatcher (Empidonax traillii extimus; SWWF), lesser long-nosed bat (Leptonycteris yerbabuenae; LLNB), and Chiricahua leopard frog (Rana chiricahuensis). A more detailed discussion of these species is provided below.

In addition, the May 10, 2011, WildEarth Guardians v. Salazar court settlement requires the USFWS to submit a Proposed Rule or a not-warranted finding to the Federal Register for over 250 species currently considered as "candidates" for listing under the endangered species act. On that list is one species, the Sonoran desert tortoise (Gopherus morafkai), which may occur in the project area or vicinity. The USFWS is required to make a determination regarding the Sonoran desert tortoise no later than the end of FY 2015 (September 30, 2015). For that reason, a more detailed discussion regarding that species is provided here.

Southwestern Willow Flycatcher

The southwestern willow flycatcher was listed as "endangered" by the USFWS in 1995, following candidacy in 1989 and proposed listing in 1994. Critical habitat was originally established in 1997 but set aside in 2001. The USFWS proposed critical habitat again in October 2004, and final critical habitat was designated in October 2005. The bird is also listed as Wildlife of Special Concern by the AGFD.

The SWWF is a migratory bird that summers in southern California, southern Nevada, southern Utah, Arizona, New Mexico, western Texas, southwestern Colorado, and northwestern Mexico. Suitable breeding habitat is typically comprised of stands of dense cottonwood, willow, and tamarisk in riparian areas, with open water usually present.

Habitat for SWWF in the project vicinity is expected to be limited to dense riparian stands along the San Pedro River. The nearest SWWF critical habitat to the project area is along the San Pedro River, at the southwestern end of both Copper Creek and Bunker Hill Roads. Although impacts to critical habitat from the project are not anticipated, improvements to the access roads have the potential to affect this species. As such, survey for SWWF, based on protocols adopted by the USFWS in May 1997 (revised May 2000), may be recommended. The survey protocols fulfill the primary objectives of documenting the presence or absence of SWWF and determining their status (territorial vs. migrant). The USFWS recommends five surveys per year for project-related surveys, with the first survey to be conducted between May 15 and May 31, the second between June 1 and June 21, and the third, fourth, and fifth between June 22 and July 17. Surveys conducted in different survey periods, and multiple surveys within the third survey period, must be at least five days apart.

Survey should be completed prior to any impacts in the vicinity of the San Pedro River. From previous permitting efforts in the surrounding area, WestLand understands that SWWF nesting has occurred along this general reach of the river. As such, USFWS


SGS FORM No. A263a-2/14/13 !


may not require survey, and will initiate Section 7 consultation absent project-specific surveys.

Lesser Long-Nosed Bat

The LLNB was proposed for listing as “endangered” by the USFWS in 1987 (52 FR 25171), with the final ruling coming in 1988 (53 FR 38456), both without critical habitat. It is also an AGFD Wildlife Species of Special Concern.

The project area is within the late summer range of this species, and some forage species (i.e. agave) were observed during the site visit. In addition, potential roost sites are available in the variety of shafts and adits within the area.

No survey protocol for LLNB exists. In general, potential roost sites can be inspected for individuals and to identify guano (bat droppings), which is indicative of nectar-feeding versus insect bats. A cursory evaluation of onsite mining features (representing potential LLNB roosting habitat) identified no guano from nectivorous bats, though a more thorough evaluation will be completed well in advance of activities which could affect the features.

Chiricahua Leopard Frog

The Chiricahua leopard frog is designated as a threatened species throughout its range by the USFWS, pursuant to the Endangered Species Act of 1973 as amended (67 FR40790, June 13, 2002). Critical habitat was designated in March 2012 and includes portions of the Galiuro Mountains (Recovery Unit 4) outside of the project area. According to the listing summary, this species is absent from more than 75 percent of its known historical sites and former range. Major threats to the population status of this species include predation by introduced bullfrogs, crayfish, and predatory fish, chytrid fungus infection, habitat fragmentation, water pollution, major water manipulations, and intensive grazing.

The Chiricahua leopard frog is considered an aquatic habitat generalist. The primary vegetation types where this species is found include oak, mixed oak, and pine woodlands, but they may also be found in chaparral, grassland, and even desert. This species breeds in a variety of conditions that include natural and manmade aquatic systems. Rock-bound pools, river-overflow pools, oxbows, permanent springs, permanent pools in intermittent streams, and beaver ponds are the natural aquatic habitats where the Chiricahua leopard frog occurs. Potentially suitable man-made structures include earthen stock tanks, livestock drinkers, irrigation sloughs, wells, mine adits, abandoned swimming pools, and ornamental backyard ponds. This species is present in central and eastern Arizona at an elevation range of 3,500 to 8,000 ft and at the Arizona-Mexico border at 1,200 to 4,000 ft.

The draft Recovery Plan indicates that populations of this species were identified in a number of stock tanks within the Galiuro Mountains in the 1990’s. The most recent studies indicate that most of these populations have disappeared, potentially due to


SGS FORM No. A263a-2/14/13 !


drought conditions. The remaining population known in the Galiuros is within a perennial stream which may have some higher resistance to drought than the stock tanks.

Aquatic habitats within the project area will be evaluated for the potential to support this species.

Sonoran Desert Tortoise

The desert tortoise in the United States is split into two populations: the Mojave desert tortoise (north and west of the Colorado River) and the Sonoran desert tortoise (limited in the United States to Arizona south and east of the Colorado River). Murphy et al. (2011) have proposed the species name Gopherus morafkai for the Sonoran population. The USFWS list of Special-status Species for Pinal County identifies the Sonoran desert tortoise as G. agassizii, but both its General Species Information (USFWS 2012a) and its profile page (USFWS 2012b) refer to the Sonoran desert tortoise as G. morafkai.

The Mojave desert tortoise was listed as threatened in 1990 (USFWS 1990a). The Sonoran desert tortoise is currently a candidate for listing. It was petitioned for listing as a Distinct Population Segment with critical habitat on October 9, 2008 (WildEarth Guardians and Western Watersheds Project 2008). On August 28, 2009, the USFWS published its 90-day finding announcing that they were initiating a status review for the Sonoran desert tortoise (USFWS 2009). On December 14, 2010, the USFWS announced that the listing of the Sonoran desert tortoise was warranted but precluded by higher priority listing actions (USFWS 2010). The May 10, 2011, WildEarth Guardians v. Salazar court settlement requires the USFWS to submit a Proposed Rule or a not-warranted finding to the Federal Register for the Sonoran desert tortoise no later than the end of FY 2015 (September 30, 2015). The Sonoran desert tortoise is also designated Arizona State Wildlife of Special Concern (AGFD 1996), Forest Service Sensitive (USDA, FS Region 3 2007), and BLM sensitive (USDI, BLM 2010). It is not afforded protection under the ESA, but is considered in this evaluation due to the potential for up listing.

The USFWS identified a wide variety of threats to the Sonoran desert tortoise in its finding that listing is warranted, including habitat modification and destruction, illegal collection, domestic dog and human depredation, vehicle strikes, and inadequate regulatory mechanisms (USFWS 2010).

The Copper Creek project area is within the Sonoran desert tortoise’s known habitat and elevation range. BLM mapping depicts Sonoran desert tortoise habitat covering essentially the entire area of the access road from the San Pedro River, but only within the extreme western fringe of the mine site itself. Desert tortoises are known to occur along the lower elevations of the Galiuro Mountains approximately 6 miles southeast of the project area (Hart et al. 1992, Woodman et al. 1999). Arizona upland vegetation with rocky slopes, where desert tortoises are typically found, is present within the project area, mainly along the proposed haul road at the lower elevations of the western end. Areas near the proposed mine site are dominated by semidesert


SGS FORM No. A263a-2/14/13 !


grassland, grading into interior chaparral, and are less likely to be occupied by desert tortoises.

An analysis for the potential occurrence of Sonoran desert tortoise will be conducted prior to initiation of mine development activities.

National Historic Preservation Act Compliance

As with the ESA, federal permitting authorities are required to evaluate the potential for any proposed project to adversely impact properties listed, or eligible for listing, on the National Register of Historic Places (NRHP). Typically, a Class III (pedestrian) cultural resources survey is completed within the proposed project area to identify any register-eligible properties. If a proposed project may adversely affect such a property, under Section 106 of the NHPA, the lead federal agency must consult with the State Historic Preservation Officer and selected Native American tribes. Mitigation for most impacts is typically accomplished through the development and implementation of a Historic Properties Treatment Plan, which must be reviewed and approved by all consulting entities.

A large portion of the project area has been surveyed for cultural resources. A number of the historic mining features within the project area have been identified by archaeological consultants as eligible for listing, either separately or as a single site. Any adverse effects to these features would likely require consultation and, likely, mitigation.

Aquifer Protection Permit

In general, an Aquifer Protect Permit (APP) will be required for any activities or facilities that may result in a discharge of a contaminant to an aquifer, including tailings and waste rock facilities, and non-stormwater impoundments. A review of the APP permitting materials for AMT’s project indicates that AMT was expecting to encounter some acid-generating waste rock that would need to be temporarily stockpiled on the surface. They had agreed to routinely sample, segregate and special-handle this material, to stockpile it on a constructed foundation, and then return it to the excavation site prior to 5 elapsed years. In addition, runoff diversion ditches associated with waste rock stockpiles were required to be lined.

The APP process can be time and labor intensive, and for mining projects generally requires extensive hydrologic characterization, engineering (i.e. BADCT demonstration), geochemical characterization, and reclamation design. Completion of the APP process can take up to two years, or more.

ASLD Right-of-Way Access

In 2011, Redhawk acquired right-of-way access across lands managed by the Arizona State Land Department (ASLD) in support of the development of the exploration decline near a minor tributary of Saloon Gulch. The right-of-way includes the majority of


SGS FORM No. A263a-2/14/13 !


Copper Creek Road as well as the construction of a new all-weather access road that avoids Copper Creek itself.

For the larger mine, it is assumed that the access roads will need to be upgraded beyond what is considered in the current ASLD right-of-way in order to accommodate the anticipated increase in the volume and type of traffic required for the mining effort. In that case, a revision to the right-of-way will be required through coordination with ASLD. As part of the right-of-way application process, the following will need to be conducted:

▪ Native plant inventory - In essence, ASLD requires a “stumpage fee” for impacts to native vegetation through ASLD lands. This task includes a complete inventory of all native woody and succulent species that will be impacted by the proposed project. The state uses this inventory to calculate the amount of the stumpage fee.

▪ Cultural resources survey - A full pedestrian (Class III) cultural resources survey is required in those areas that will be impacted on state lands (see below).

The results of the native plant inventory and the cultural resources survey are submitted to ASLD for processing along with the applicable Use or Right-of-Way permit application. Although the inventory and survey are relatively straightforward efforts, the processing time can be significant, as long as a year. If cultural resources will be adversely affected by the proposed project the process will likely take longer.

State Reclamation Plan

The State of Arizona requires that mining projects greater than five (5) contiguous acres on private lands submit a reclamation plan to the State Mine Inspectors Office. This plan must outline planned reclamation activities and costs. The reclamation requirements under this program are different than those for the APP program, described above.


SGS FORM No. A263a-2/14/13 !


Dam Safety Permit

The Arizona Department of Water Resources (ADWR) has jurisdiction over non-federal dams in Arizona over 25 ft in height or capable of storing more than 50 acre-ft of water, including appurtenances. Smaller dams, or those within certain defined classifications, are exempt. For new dams, or those requiring alteration or repair, application to and approval from the Dam Safety Section of ADWR is required. Applicants are encouraged to meet with ADWR staff prior to application submittal.

Tailings dams are specifically exempt from permitting under this program. It is assumed for the purpose of this analysis that there will be no other jurisdictional dams as part of this project.

Storm Water Permit (MSGP)

Under the Arizona Pollutant Discharge Elimination System (AZPDES) Multi-Sector General Permit (MSGP-2010) Storm water program administered by the Arizona Department of Environmental Quality (ADEQ), mine facilities (including associated pre-mining exploration) are required to obtain coverage for discharges of storm water from their operations. This program requires a project proponent to prepare a storm water pollution prevention plan (SWPPP), submit a notice of intent (NOI) to discharge storm water, install appropriate best management practices (BMPs), and conduct regular inspections and of the site and analytical monitoring during construction and operations, in accordance with the SWPPP.

The Copper Creek operations are currently covered under the MSGP for the historic mining operations within the project area.

Individual AZPDES Permit

Discharges of certain industrial process waters are covered by individual permits under the AZPDES program. Anticipated process water discharges from a mine site include water treatment plant discharges, wastewater from mill operations, and seepage from tailings impoundments.

An issued AZPDES permit includes effluent limitations, usually consisting of both numeric and narrative standards. The numeric limitations typically restrict quantities, rates, and concentrations of pollutants that may be present in the discharge, and can be either technology- or water quality-based. Technology-based standards require usage of available pollution control technology, while water quality-based standards protect ambient water quality by requiring the discharger to achieve the applicable numeric standard (as mentioned above). If both technology- and water quality-based standards exist for a particular constituent, the more restrictive standard applies.

It is not anticipated that the proposed project will not require coverage under an individual AZPDES permit.


SGS FORM No. A263a-2/14/13 !


Permitting Considerations

Critical path items related to environmental permitting for the project are:

▪ BLM Approvals (including NEPA, ESA, and NHPA compliance), ▪ Clean Air Act permitting ▪ CWA Section 404 permit (including NEPA, ESA, and NHPA compliance), and ▪ Aquifer Protection Permit

In general, other approvals are readily acquired within the critical path. A summary of environmental permitting requirements is provided in the table below.

Table 20.2 Environmental Permitting Analysis for Copper Creek Project

Permit Effort AgencyDescription/

Assumptions Estimated Cost*

Estimated Timefram

eSchedule

Start Point

MPO/NEPA Bureau of Land Management (BLM)

Assumes that level of impacts will require an Environmental Impact Statement (EIS)

1,500,000 2 to 5 years

Availability of location and operating information for federal land facilities

Air permit ADEQ/Pinal County

Title V permit required for the mill with new source review (NSR), and Prevention of Significant Deterioration (PSD) requirements; includes some modeling

250,000 24 to 36 months

Availability of detailed emissions and ambient air information, submittal of application

CWA Section 404/NEPA

U.S. Army Corps of Engineers

For all discharges of fill to waters; assumes an individual permit will be required for tailings facilities and reroute of Copper Creek

200,000 18 to 24 months

Completion of JD; submittal of application; preparation of mitigation plan, alternatives analysis, etc

Endangered Species Act Compliance

Lead federal agency

Required for all federal actions; assumes informal consultation for potential impacts to one or more species

40,000 6 months Availability of locations of operations and Phase I design information

National Historic Preservation Act Compliance

Lead federal agency and SHPO

Includes Class I and Class III survey, treatment plan, and coordination. Data recovery not included

120,000 12 months Availability of disturbance boundary

Aquifer Protection Permit

ADEQ APP needed for waste rock and tailings facilities; monitoring well installation required

60,000 12 to 15 months

Submittal of application with baseline data and basic engineering


SGS FORM No. A263a-2/14/13 !


*US$; permitting costs only; does not include engineering, well installation, etc.

Right-of-Way Access

Arizona State Land Department

Assumes that roadway widening or other significant modification will be required for access; includes resource surveys

25,000 12 months Availability of road designs

Reclamation Plan

State Mine Inspector

Needed for mining disturbances over 5 acres on private land.

15,000 3 months Availability of general arrangement, and geotechnical and geochemical information

Dam Safety Permit

ADWR Needed for jurisdictional impoundments (greater that 25 ft embankment height or greater than 50 ac-ft storage capacity); assumed not required for the Copper Creek Project.

NA NA NA

Stormwater General Permit

ADEQ Triggered by having qualifying stormwater discharges to Waters of the US; requires updating SWPPP and submitting revised NOI

25,000 2 months Availability of general arrangement

Total 2,230,000+ Up to 5 years

Table 20.2 Environmental Permitting Analysis for Copper Creek Project

Permit Effort AgencyDescription/

Assumptions Estimated Cost*

Estimated Timefram

eSchedule

Start Point


SGS FORM No. A263a-2/14/13 !


21.0 CAPITAL AND OPERATING COSTS

21.1 Underground. Mine

21.1.1 Underground Mine Capital Estimates

The mine capital required to develop and sustain a 25,000 t/d Post-Pillar Cut & Fill mining operation for extracting Copper Creek materials from the Childs, Mammoth, Keel, and American Eagle deposits has been estimated to a plus or minus (+/-) 35 percent accuracy. These costs are presented in first quarter 2013 U.S. dollars.

The costs are based on unit prices applied to estimated development quantities, recent equipment quotes from mining equipment manufactures, and labor cost build-ups for the mine site location, similar installations, and in-house information on labor productivities for large-tonnage underground room and pillar operations.

The following Table 21.1 provides a summary breakdown of the mining capital required to develop and sustain a 25,000 ton per day, fully integrated mining and processing facility at Copper Creek, over an average 18 year mine life.

Table 21.1 Mine Capital Cost Summary

Cost Center Preproduction Capital

Sustaining Capital

Total Capital

Mine:

Breccia Decline $5,040,000 $0 $5,040,000

Shaft Site Prep 3,000,000 0 3,000,000

Prod. & Vent Shafts 163,680,000 0 163,680,000

Level Development 64,283,000 0 64,283,000

Vent/Rock Passes 12,000,000 25,500,000 37,500,000

Internal Ramping 0 40,800,000 40,800,000

Miscellaneous Small AFE’s 2,000,000 8,500,000 10,500,000

Equipment 118,161,000 73,880,000 191,241,000

Subtotal Mine $368,840,000 $148,680,000 $517,544,000

Mine Freight, Commissioning $9,509,000 $5,910,000 $15,419,000

Mine EPCM $26,486,000 $0 $26,486,000

Preproduction-. Owner’s Costs: $ 12,000,000 $0 $12,000,000


SGS FORM No. A263a-2/14/13 !


Capex before Contingency $416,860,000 $154,590,000 $571,450,000

Contingency @ 30% $125,058,000 $46,377,000 $171,435,000

Total Estimated Mine Capital $ 541,918,000 $200,967,000 $742,885,000


SGS FORM No. A263a-2/14/13 !


The estimated annual mine capital expenditures, without contingency, are shown in the following Table 21.2.

21.1.2 Hydraulic Backfill Cost Estimate

Hydraulic backfill is the preferred method for backfill at the Copper Creek Mine in this Preliminary Economic Assessment (PEA). Hydraulic backfill is slurry typically

Table 21.2 Estimated Annual Capital

Year Estimated Expenditure

-4 $57,998,000

-3 46,774,000

-2 113,007,000

-1 199,081,000

Subtotal Preproduction $416,860,000

1 $4,400,000

2 4,400,000

3 4,400,000

4 4,400,000

5 4,400,000

6 31,400,000

7 4,400,000

8 4,400,000

9 4,400,000

10 31,400,000

11 4,400,000

12 4,400,000

13 4400,000

14 30,190,000

15 4,400,000

16 4,400,000

17 4,400,000

18 0

Subtotal Sustaining $154,450,000

Total Mine Capital $571,450,000


SGS FORM No. A263a-2/14/13 !


consisting of coarse tailings particles and a binder (such as cement). The solid content of typical hydraulic backfill ranges from 65 to 72 weight percent solids. Binder is added to the slurry in proportion to the weight of solids in the slurry. Hydraulic backfill will generally require more binder content to achieve the same backfill strength as a typical paste backfill. It is often observed that the Unconfined Compressive Strength (UCS) requirement for stopes may vary depending on the mining method and the location of the stopes.

Scoping level capital and operating cost estimates, with an intended overall level of accuracy of ±40 percent, have been established for sourcing, loading, hauling and sizing the required backfill make-up, and operating an aggregate/sand plant and Hydraulic Backfill Plant. The following assumptions and basis were used to generate the cost estimate:

▪ All costs are in first quarter 2013 US dollars;

▪ A base date of January 2013;

▪ Current costs associated with items such as contractor and installation labor and consumable items e.g. power and cement, etc., are based on Golder’s database for similar projects;

▪ Costs for the majority of the process equipment were established using actual historical pricing for similar equipment, as close in capacity as possible to the ones considered for the Project.

▪ All equipment costs are based on new purchased equipment;

▪ The TSF thickener is not included in the scope of the backfill plant;

▪ Since the in situ backfill strength requirement is unknown at the time of the study, backfill cement consumption rate for the cemented hydraulic backfill was estimated based on Golder’s experience in other mine backfill applications.

▪ Binder (NPC) is available for delivery to the Copper Creek mine site at $100/ton delivered.

▪ The labor costs for all capital activities is predicated around the work being performed by qualified contactors; all labor rates are presented in 2013 dollars and estimated to be $80 per hour;

▪ The breakdown of plant direct costs is factored estimates based on Golder’s database of previous projects; and

▪ Indirect costs were not included in the capital cost estimates.


SGS FORM No. A263a-2/14/13 !


Based on the above criteria and assumptions, the estimated capital and operating cost estimates for providing make-up backfill sand, and operating the Aggregate and the Hydraulic Backfill Plants are summarized in Tables 21.3 and 21.4 below. The capital of the aggregate/sand plant is assumed to be leased and expensed as an operating cost. The spare capacity of the aggregate plant may be used to produce crushed stone for on-site general purpose material such as concrete, ballast, and road building

21.1.3 Discussion

Binder represents the largest portion of the backfill plant operating cost, and will be subjected to extensive testing of various binder combinations, ratios, and admixtures to optimize the cost per unit strength benefit. Typically binder percentages range from 4 to 12 percent of the slurry solids by weight and can represent 80 to 90 percent of the cost of backfill

Table 21.3 Hydraulic Backfill Plant

Capital Cost Estimate (±40%)

Description Capital Cost Values, $USD

Process Mechanical Costs 5,757,000

Direct Plant Costs 5,800,000

Total Capital Costs 11,557,000

Table 21.4 Hydraulic Backfill Plant

Operating Annual Cost Estimate (±40%)

Description Operating Cost Values, $USD

Consumables

Power 743,000

Binder 18,729,000

Other 402,000

Hyd. Backfill Plant Surface Labor 635,000

Manufactured Sand Plant (lease) 1,080,000

Operate Sand Plant 2,790,000

Load/Haul/Stack Coarse 21,611,000

Total Operating Costs 45,990,000

Equivalent cost per ton of ore 5.11


SGS FORM No. A263a-2/14/13 !


21.2 Underground Mine and G&A Operating Estimates

Mine operating costs have been estimated, based on producing 25,000 t/d, using a post-pillar cut and fill mining method, from four separate deposits, simultaneously. The estimate was based on prevailing wage rates in the U.S copper mining industry, and underground equipment operating costs in average ground conditions. An electric power cost of US$ 0.077/kWh was used for estimating power costs. All costs are in first quarter 2013 US dollars. A summary of this estimate for full production years 4 through 18 is provided in the following Table 21.5.

Table 21.6 provides a summary of the estimated annual mine operating costs. Tables 21.7 and 21.8 provide a breakdown of the operating labor cost build-up and a summary of the full production mine average unit operating cost estimate. Table 21.9 lists the estimated full production G&A operating costs.

A reduction of 20 percent in tonnage, and a corresponding increase in operating costs, was included for the learning curve in the first year of production. In addition, the first three production years reflect the higher operating cost of mining the upper breccia material. The production ramp-up is expected to be very steep, since the preproduction development work allows immediate access to the mineralized material. From this initial access, mining can begin with multiple headings.

Table 21.5 Full Production

Mine Operating Cost Summary

Labor, $/t Supplies, $/t Total, $/t

Direct Costs

Production 4.15 8.59 12.74

Development 0.32 0.27 0.59

Service & Support 1.30 2.45 3.75

Subtotal Direct 5.77 11.31 17.08

Indirect Costs

Supervision/Technical 0.31 0.06 0.38

Subtotal Indirect 0.31 0.06 0.38

Total Mine 6.09 11.37 17.45

Table 21.6 Estimated U.G. Mine Annual Operating Expense


SGS FORM No. A263a-2/14/13 !


Year Unit Cost, $/stAnnual

Expenditure, $

1 19.89 143,208,000

2 19.05 171,450,000

3 18.25 164,250,000

4 17.45 157,050,000

5 17.45 157,050,000

6 17.45 157,050,000

7 17.45 157,050,000

8 17.45 157,050,000

9 17.45 157,050,000

10 17.45 157,050,000

11 17.45 157,050,000

12 17.45 157,050,000

13 17.45 157,050,000

14 17.45 157,050,000

15 17.45 157,050,000

16 17.45 157,050,000

17 17.45 157,050,000

18 17.45 81,841,000

Total Production 17.70 2,759,449,000


SGS FORM No. A263a-2/14/13 !


Table 21.7

Labor Basis Copper Creek Post-Pillar Cut & Fill PEA Study

Operating Labor Cost Estimate 25,000 tpd Option

No. Labor ClassificationBase $/

Hr Incentive Fringe @ 40%O'tim

e Total $/Hr Cost/MS Cost/day

Direct:

54 Miner $30.00 $15.00 $12.00 $2.10 $59.10 $472.80 $25,531.20

54 Miner Helpers $28.00 $14.00 $11.20 $1.96 $55.16 $441.28 $23,829.12

116 Equipment Operators $28.00 $14.00 $11.20 $1.96 $55.16 $441.28 $51,188.48

24 Backfill Operators $26.00 $13.00 $10.40 $1.82 $51.22 $409.76 $9,834.24

3 Hoistmen $32.00 $13.00 $12.80 $2.24 $60.04 $480.32 $1,440.96

3 Toplander $22.00 $13.00 $8.80 $1.54 $45.34 $362.72 $1,088.16

20 Mechanic $28.00 $0.00 $11.20 $1.96 $41.16 $329.28 $6,585.60

20 Mechanic Helpers $24.00 $0.00 $9.60 $1.68 $35.28 $282.24 $5,644.80

5 Electrician $30.00 $0.00 $12.00 $2.10 $44.10 $352.80 $1,764.00

4 Electrician Helpers $26.00 $0.00 $10.40 $1.82 $38.22 $305.76 $1,223.04

8 Definition Drillers $28.00 $0.00 $11.20 $1.96 $41.16 $329.28 $2,634.24

36 Nipper $22.00 $0.00 $8.80 $1.54 $32.34 $258.72 $9,313.92

21 Labor $18.00 $0.00 $7.20 $1.26 $26.46 $211.68 $4,445.28

368 Total Direct Labor $144,523.04

Cost/ton $5.78

*Scheduled (7-day rotation@ 6.35%) & Unscheduled (1.65%) Overtime = 8%

No. Indirect: Fringe @ 30% Total Cost Cost/Year

1 Mine Manager $150,000 $45,000 $195,000 $195,000

3 General Foremen $85,000 $25,500 $110,500 $331,500

9 Shift Bosses $70,000 $21,000 $91,000 $819,000

1 Chief Engineer $85,000 $25,500 $110,500 $110,500

3 Engineers/Planners $60,000 $18,000 $78,000 $234,000

2 Surveyors $45,000 $13,500 $58,500 $117,000

1 Chief Geologist $85,000 $25,500 $110,500 $110,500

3 Geologists $60,000 $18,000 $78,000 $234,000

6 Samplers $40,000 $12,000 $52,000 $312,000

3 Mine Office $50,000 $15,000 $65,000 $195,000

3 Mine Dry $40,000 $12,000 $52,000 $156,000

35 Subtotal Indirect $770,000 $231,000 $2,814,500

Cost/ton $0.31

403 Total Mine Unit Labor Cost $6.09


SGS FORM No. A263a-2/14/13 !



SGS FORM No. A263a-2/14/13 !


Table 21.8 Copper Creek Post-Pillar Cut & Fill PEA Study

Mine U.G. Operating Cost Estimate 25,000 tpd Option

No. Cost Center Labor, $/t Materials, $/t Subcontractors Total, $/t

U.G. DIRECT

Production

90 Drilling & Blasting $1.65 1.48 $0.00 $3.13

60 LHD Mucking $1.06 1.09 $0.00 $2.15

48 U.G. Haulage $0.85 0.85 $0.00 $1.70

24 Backfilling $0.39 5.11 $0.00 $5.50

6 Hoisting $0.10 0.02 $0.00 $0.12

6 Surface Haulage $0.11 0.04 $0.00 $0.15

234 Subtotal Production $4.15 $8.59 $0.00 $12.74

Development

6 Ramping $0.11 0.10 $0.00 $0.21

6 Level Development $0.11 0.09 $0.00 $0.20

8 Definition Drilling $0.11 0.08 $0.00 $0.19

20 Subtotal Development $0.32 $0.27 $0.00 $0.59

Service & Support

6 Ground Repair $0.11 0.12 $0.00 $0.23

6 Ventilation $0.06 0.02 $0.00 $0.08

6 Pumping $0.06 0.03 $0.00 $0.09

3 Sanitation $0.03 0.01 $0.00 $0.04

2 Roads $0.03 0.02 $0.00 $0.05

40 Mechanical Maintenance. $0.49 0.68 $0.00 $1.17

9 Electrical Maintenance. $0.12 0.06 $0.00 $0.18

0 Electric Power $0.00 1.45 $0.00 $1.45

24 Materials Handling $0.25 0.02 $0.00 $0.27

18 U.G. Labor $0.15 0.03 $0.00 $0.18

114 Subtotal Serv. & Support $1.30 $2.45 $0.00 $3.75

368 Subtotal Direct $5.77 $11.31 $0.00 $17.08

U.G. INDIRECT


SGS FORM No. A263a-2/14/13 !


Supervision/Technical

1 Mine Manager $0.02 0.00 $0.00 $0.03

3 General Foremen $0.04 0.01 $0.00 $0.04

9 Shift Bosses $0.09 0.02 $0.00 $0.11

1 Chief Engineer $0.01 0.00 $0.00 $0.01

3 Engineers/Planners $0.03 0.01 $0.00 $0.03

2 Surveyors $0.01 0.00 $0.00 $0.02

1 Chief Geologist $0.01 0.00 $0.00 $0.01

3 Geologists $0.03 0.01 $0.00 $0.03

6 Samplers $0.03 0.01 $0.00 $0.04

3 Mine Office $0.02 0.00 $0.00 $0.03

3 Mine Dry $0.02 0.00 $0.00 $0.02

35 Subtotal Indirect $0.31 $0.06 $0.00 $0.38

403 TOTAL U.G. MINE $6.09 $11.37 $0.00 $17.45


SGS FORM No. A263a-2/14/13 !


Table 21.9 G&A Costs

25,000 tpd Option

No. Classification Base Salary Fringe @ 35% Total/Year

Labor

1 General Manager $200,000 $70,000 $270,000

1 Assistant Manager $130,000 $45,500 $175,500

1 Administration Manager $80,000 $28,000 $108,000

1 Environmental Manager $90,000 $31,500 $121,500

1 Warehouse Manager $75,000 $26,250 $101,250

1 Purchasing Manager $75,000 $26,250 $101,250

1 Safety Manager $75,000 $26,250 $101,250

1 Technical Services Mgr. $90,000 $31,500 $121,500

1 Maintenance Manager $110,000 $38,500 $148,500

8 Secretaries $40,000 $14,000 $432,000

12 Clerks $40,000 $14,000 $648,000

5 Accountants $55,000 $19,250 $371,250

2 Safety/Trainers $50,000 $17,500 $135,000

4 Warehousemen $45,000 $15,750 $243,000

2 Purchasers $45,000 $15,750 $121,500

3 Engineers/Quality Control $70,000 $24,500 $283,500

2 Maintenance Planners $70,000 $24,500 $189,000

1 Road Maintenance $55,000 $19,250 $74,250

3 Infrastructure Maintenance $45,000 $15,750 $182,250

6 Surface Equip. Maintenance $65,000 $22,750 $526,500

57 Total Labor $4,455,000

Cost Per Ton $0.50

Supplies Cost/Month Annual Total

Utilities $20,000 $240,000

Insurance/Taxes $17,000 $204,000

Donations $3,000 $36,000

Infrastructure Maintenance $8,500 $102,000

Road Maintenance $5,000 $60,000

Office Supplies $5,000 $60,000

Safety/Training Supplies $3,000 $36,000

Environmental $1,000 $12,000


SGS FORM No. A263a-2/14/13 !


VIP Travel $7,000 $84,000

G&A Equip. Maintenance $8,000 $96,000

Misc. Supplies $1,000 $12,000

Total Supplies $942,000

Cost/Ton $0.10

Total Unit cost $0.60


SGS FORM No. A263a-2/14/13 !


21.3 Processing Plant

21.3.1 Process Capital Costs

A summary of the initial capital costs is shown in the following Table 21.10. The table includes direct costs, indirect costs, and a contingency. A detailed description of the estimate is presented in Appendix 8.

Table 21.10 Summary Process Plant - Capital Cost for Base Case 25,000 TPD

Description Total Cost, US$

DIRECT COSTS

Area 10 – Primary Crushing 10,000,477

Area 15 – Grinding 37,372,924

Area 40 – Copper Moly Flotation 12,174,266

Area 45 – Moly Flotation 3,548,636

Area 50 – Copper Concentrate Handling 2,327,903

Area 55 – Moly Concentrate Handling 3,239,561

Area 60 – Reagents 2,497,033

Area 70 – Tailing Handling 2,005,661

Area 80 – Reclaim and Process Water 1,031,355

Area 90 – Fresh Water 661,312

SUB-TOTAL INSTALLED EQUIPMENT COST 74,859,128

Process Piping 8,020,621

Electrical 13,367,702

Instrumentation 8,020,621

Site Development 10,694,161

Buildings 13,367,702

Roads (Paved, 10.6 miles) 25,440,000

Overhead Transmission Line 16,041,242

Utilities and Outside Lines 8,020,621

SUB-TOTAL DIRECT 177,831,797

INDIRECT COSTS

EPCM 10,694,161

Construction Indirect Costs include: 17,783,180

Construction Supervision Included

Equipment Rental Included


SGS FORM No. A263a-2/14/13 !


Field Office Expense Included

Mobilization / Demobilization Included

Consumables Included

Owner’s Costs 5,347,081

Spare Parts 2,673,540

Initial Fill & Reagents 2,667,477

Equipment Insurance & Freight Cost 4,277,664

SUB-TOTAL INDIRECT 43,443,103

TOTAL DIRECT AND INDIRECT 221,274,900

Contingency – 35% 77,446,215

TOTAL COST 298,721,115


SGS FORM No. A263a-2/14/13 !


Direct Costs

The direct capital costs were based on the following list of documents:

▪ Design Criteria ▪ Equipment List ▪ Mining Cost Service Source Quote Data ▪ SGS Equipment Database ▪ Engineering Drawings performed by SGS

- 10-F-01 Flowsheet & Mass Balance Primary Crusher - 15-F-02 Flowsheet & Mass Balance Grinding - 40-F-06 Flowsheet & Mass Balance Flotation - 40-F-07 Flowsheet & Mass Balance Regrind - 40-F-08 Flowsheet & Mass Balance Copper-Moly Flotation - 45-F-09 Flowsheet & Mass Balance Moly Flotation - 45-F-10 Flowsheet & Mass Balance Moly Cleaning - 50-F-11 Flowsheet & Mass Balance Concentrate Handling - 55-F-13 Flowsheet & Mass Balance Moly Concentrate Handling - 60-F-30 Flowsheet & Mass Balance Reagents - 60-F-31 Flowsheet & Mass Balance Reagents - 70-F-35 Flowsheet & Mass Balance Tailings Handling - 80-F-40 Flowsheet & Mass Balance Process and Reclaim Water - 90-F-45 Flowsheet & Mass Balance Raw Water - 05-G-001 Civil Project Area Overall View - 05-G-002 Civil Property Area Site Plan - 05-L-010 General Arrangement Plant Layout - 05-L-015 General Arrangement Grinding Area Layout - 05-L-020 General Arrangement Flotation and Thickener Area Layout

The direct costs in this estimate include, but are not limited to, labor, equipment and materials for the detailed construction activities set forth below:

▪ Equipment Costs

An equipment list was developed and incorporated into the cost estimate. The estimate for equipment was developed from the following sources:

- Written or e-mailed budgetary estimates from vendors for major equipment.

- Historical data and budget costs from similar projects for miscellaneous equipment.

The cost for “Installed Equipment” was estimated using a factor of forty percent (40%) of purchased equipment costs. This factor reflects typical costs to install equipment and covers labor, concrete, steel, and other services and construction materials associated with equipment erection and placement.


SGS FORM No. A263a-2/14/13 !


▪ Process Piping

Process piping costs include materials and installation of all piping within the process plant. The process piping cost was estimated using a fifteen percent (15%) factor of purchased equipment costs. The factor was selected based on the size of the plant, as well as the distribution of piping within the plant area.

▪ Electrical Distribution and Instrumentation

Electrical and Instrumentation costs include transformation and service, wiring, cable tray, lighting and grounding within the process plant. The electrical and instrumentation cost was estimated using twenty-five percent (25%) and fifteen percent (15%) factors, respectively, of purchased equipment costs. The factors were selected based on preliminary equipment power requirements and latest National Electric Code (NEC) standards.

▪ Site Development

Site development costs include excavations, backfills, grading, roads, and fencing. The construction site development cost was estimated using a twenty percent (20%) factor of the purchased plant equipment cost. The factor was selected based on a rough civil computer model to calculate approximate bulk excavations and backfills based on the current plant layout of the proposed project site. The Project will require development at the following major locations:

- New access roads - Process plant area - Crushing and conveying areas

▪ Buildings

Building costs include materials, labor, and other miscellaneous costs associated with erecting covered structures within the project site. The building costs were estimated using a twenty-five percent (25%) factor of the purchased plant equipment cost. The factor was selected to reflect the projected costs of the buildings based on building type and square footage. The Project will require the following buildings:

- Mill Building - Flotation Building - Cu Concentrate Load Out Building


SGS FORM No. A263a-2/14/13 !


▪ Roads

Road costs include materials, labor and other miscellaneous costs associated with constructing new roads and upgrading existing ones. Historically, costs for roads have typically been approximately $2.4M USD per mile. These observed costs have been utilized in this estimate for the 10.6 miles of paved roads to the project site.

▪ Overhead Transmission Line

The costs for supplying power to the Project includes a utility transmission line including cables, poles, conductors and insulators, labor and other miscellaneous costs associated with erecting the utility transmission line to site and costs for on-site power distribution that includes transformers, switchgear, labor and other miscellaneous costs. The cost was estimated using a thirty percent (30%) factor of purchased plant equipment cost. The factor was selected to reflect a cost allowance for a conceptual plan to have power supplied to the Copper Creek Mine main substation via a 115 kV transmission line from Arizona Public Service (APS) substation, at San Manuel, and distribute power on site via transformers and switchgear to all project facilities with overhead and/or underground power lines. The Project will require the following main electrical power supply and distribution items:

- Utility Transmission Line

- Main Transformer(s) stepping down utility transmission line voltage to on-site distribution line voltage.

- Main switchgear for distributing power on site

- On-site power lines, overhead and/or underground

- Transformation and switchgear as needed for service at each respective facility of the project

- Backup power generators

▪ Utilities and Outside Lines

Utilities and outside lines costs include the supply of fresh water, reclaim water, fire protection, air, fuel, sanitary systems, and other miscellaneous service items within the project site. The site development cost was estimated using a fifteen percent (15%) factor of the purchased plant equipment cost. The factor was selected to reflect the projected costs of the pipelines for fresh and reclaim water transmission, with consideration of additional costs for services listed above.


SGS FORM No. A263a-2/14/13 !


Indirect Costs

Certain indirect costs exhibited in this estimate include, but are not limited to, labor, equipment and materials for the detailed activities set forth below:

▪ EPCM was estimated using a twenty percent (20%) factor of the purchased equipment costs. The factor and includes:

- Prefeasibility Engineering - Bankable Feasibility Engineering - Detailed Engineering - Procurement - Construction Management - Training

▪ Construction Indirect Costs were estimated using a ten percent (10%) factor of the total direct costs and includes:

- Construction Supervision - Equipment Rental - Field Office Expenses - Mobilization/Demobilization - Consumables

▪ Owner’s Costs were estimated using a ten percent (10%) factor of the total direct costs. These costs account for all expenditures made by the Owner during the project, including, but not limited to, management, office expenses, travel, etc. This number may need to be increased if the owner wants to include working capital in this estimate.

▪ Spare Parts costs were estimated using a five percent (5%) factor of the installed plant equipment cost.

▪ Initial Fill & Reagents costs were estimated using a one and half percent (1.5%) factor of the installed plant equipment costs.

▪ Equipment Insurance and Freight costs were estimated using an eight percent (8%) factor of the installed plant equipment costs.

Contingency and Accuracy

The SGS crushing and process plant portion of the cost estimate includes a thirty-five percent (35%) contingency for project unknowns and identified risks. Contingency is a necessary part of the cost estimate and is based on the fact less than five percent (< 5%) of the engineering is completed to date. SGS believes the estimated contingency amount will be spent during the life of the project for identified risks and unknown items.


SGS FORM No. A263a-2/14/13 !


While SGS has not performed a statistical analysis of the crushing plant and process plant accuracy of the capital cost estimate, SGS has a high confidence, based on previous experience with similar projects, that the accuracy of the process portion of the PEA capital cost estimate will end up between minus ten percent and plus thirty-five percent (-10 / +35%) of the SGS capital cost estimate.

Exclusions SGS has excluded the following cost items from the process plant estimate and they are included elsewhere:

▪ Geotechnical ▪ Mining ▪ Reclamation and Closure ▪ Metallurgical Testing ▪ Property Acquisition ▪ Permitting ▪ Environmental ▪ Permits, Royalties and Licenses ▪ Taxes, Duty and Import Fees; IVA ▪ Local Sales & Import Taxes ▪ Hazardous Waste Removal ▪ Other Consultants

Risk Sensitivity Due to minimal engineering and the utilization of a factored cost estimate, a basic risk sensitivity analysis was performed to determine a potential range of total capital cost. Although the majority of equipment costs were provided based on previous quotes and historical data for similar equipment, there remains a risk that a more detailed evaluation would result in a considerable variation in costs. SGS is confident that the equipment costs provided for this estimate are within the range of plus or minus twenty percent (+/- 20%). Performing the risk sensitivity with this range for the equipment costs results in a total capital cost range of 255M USD to 351M USD. This result is based on the assumption that all other assumptions remain valid and all factors currently selected remain unchanged.

21.3.2 Operating Costs

Annual and unit process operating cost estimates for a 25,000 ton per day milling operation are summarized in the following Table 21.11 Support tables for the cost estimates are shown in Tables 21.12 through 21.17.

Table 21.11 Summary of Plant Operating Cost by Cost Item

Item Annual Cost, US$ Cost, US$/ton


SGS FORM No. A263a-2/14/13 !


The detailed power consumption estimate is based on the installed power with estimates of the operating power draft and operating time, and power unit cost of US$ 0.077/kWh. The process power consumption and power cost calculation are summarized in Table 21.12 and Table 21.13 respectively.

Power 13,231,043 1.47

Labor 7,651,171 0.85

Reagents 4,668,350 0.52

Grinding media 10,539,310 1.17

Repair materials and operating supplies 2,138,832 0.24

Liners and wear materials 5,319,736 0.59

Water supply 1,248,114 0.14

Total 44,796,556 4.98

Table 21.12 Power Consumption Summary

Area kWh/ton

Area 10 Primary Crushing 1.975

Area 15 SAG Grinding 12.489

Area 40 Copper - Copper Moly Flotation 2.497

Area 45 Moly Flotation 0.222

Area 50 Copper Concentrate Handling 0.197

Area 55 Moly Concentrate Handing 0.216

Area 60 Reagents 0.338

Area 70 Tailing Handling 0.076

Area 80/90 - Process & Reclaim Water / Fresh Water 1.083

Total 19.09

Table 21.13 Power Cost

Usage

kWh per ton 19.09

Power Cost, US$ per kWh $0.077

Power Cost, US$ per ton $1.47

Power Cost, US$ per year $13,231,043


SGS FORM No. A263a-2/14/13 !


The labor cost estimate for mill operations is shown in Table 21.14. The labor rates and burden are based on the rates for similar mill operation.


SGS FORM No. A263a-2/14/13 !


Table 21.14 Labor Cost

Function Per Crew Total Total hrs/Yr Rate Total

Operations Shift Crews (4 crews req'd)

Control Room Operator 1 4 8,760 40.33 353,291

Grinding Operator 1 4 8,760 31.95 279,882

Copper Flotation Operator 1 4 8,760 31.95 279,882

Mo Flotation Operator 1 4 8,760 31.95 279,882

Filter / Dryer Operator 1 4 8,760 31.95 279,882

Tailings / Water Operator 1 4 8,760 31.95 279,882

Training / Vacation Relief 1 4 8,760 28.54 250,010

Sub total 28 2,002,711

Operations Day Crew

Cu conc handling (loading trucks) 2 2 4,160 28.54 118,726

Mo conc packing 1 1 2,080 28.54 59,363

Reagent mixing (10 and 4) 2 2 4,160 28.54 118,726

Tailings dam operation 6 6 12,480 28.54 356,179

General cleanup; ball charging 3 3 6,240 28.54 178,090

Sub total 12 831,085

Maintenance

Mechanics 14 14 29,120 33.47 974,646

Shift Electrician 1 4 8,760 36.71 321,580

Day Electrician / Inst. Tech 8 8 16,640 36.71 610,854

Laborers 2 2 4,160 28.54 118,726

Sub total 28 2,025,807

Technical

Shift Sample Prep / Sampler 1 4 8,760 27.6 241,776

Day sample prep 2 2 4,160 27.6 114,816

Assayers (day only) 5 5 10,400 34.24 356,096

Sub total 11 712,688

Salaried Personnel

Mill Superintendent 1 158,760 158,760

General Foreman 1 110,760 110,760

Maintenance Foreman 3 107,920 323,760

Maintenance Engineer 1 122,120 122,120

Maintenance Technician 1 82,360 82,360


SGS FORM No. A263a-2/14/13 !


Reagent cost estimates are shown in Table 21.15. The reagent consumption rates are based on METCON metallurgical test work data in 2012.

Operations Shift Foreman 4 109,770 439,080

Operations Day Foreman 1 107,920 107,920

Senior Chemist 1 115,000 115,000

Senior Metallurgist 1 122,120 122,120

Metallurgist 3 110,760 332,280

Process Technician 2 82,360 164,720

Sub total 19 2,078,880

Grand Total 100 7,651,171


SGS FORM No. A263a-2/14/13 !


The grinding media and liner and wear material cost estimates are provided in Table 21.16 and 21.17. The consumption estimates are based on abrasion index.

Table 21.15 Reagent Costs

UsageQuantity lb/year

Reagent Cost, $/lb

Cost, $/year

Cost, $/t Orekg/mt Ore lb/st Ore

Reagents

Bulk Flotation

C-3330 0.005 0.010 90,000 0.77 $69,400 $0.01

Lime 0.900 1.800 16,200,000 0.10 $1,653,344 $0.18

MIBC 0.020 0.040 360,000 1.45 $522,000 $0.06

A-238 0.003 0.006 54,000 2.38 $128,520 $0.01

NaHS 0.172 0.344 3,096,000 0.59 $1,825,619 $0.20

Fuel Oil 0.015 0.030 270,000 0.50 $134,717 $0.01

Thickening

Flocculant 0.0100 0.020 180,000 1.86 $334,751 $0.04

Total $4,668,350 $0.52


SGS FORM No. A263a-2/14/13 !


Table 21.16 Grinding Media Operating Cost Estimates

Bond Wear Equations

Usage, Pounds per

kWh

Power Consumpti

on kWh per ton

Usage, Pounds per ton

Actual Usage, Pounds per ton

Cost, $ per

pound

Cost, $ per ton

oreCost, $ per year

SAG Mill Balls 0.239 0.67 0.160$1,444,4

90

Ball Mill Balls=0.35 x (Ai - 0.015)^(1/3) 0.215 6.800 1.464 1.464 0.67 0.983

$8,850,139

Regrind Mill Balls

=0.35 x (Ai - 0.015)^(1/3) 0.215 0.188 0.040 0.040 0.67 0.027

$244,680

Total Grinding Media 1.171

$10,539,310

Table 21.17 Wear Material Operating Cost Estimates

Bond Wear Equations

Usage, Pounds per

kWh

Power Consumpti

on kWh per ton

Usage, Pounds per ton

Actual Usage, Pounds per ton

Cost, $ per

pound

Cost, $ per ton

oreCost, $ per year

SAG Mill liners 0.100 2.72 0.272$2,447,5

62

Ball Mill liners=0.026 x (Ai - 0.015)^0.3 0.017 6.800 0.114 0.114 2.72 0.311

$2,794,903

Regrind Mill liners

=0.026 x (Ai - 0.015)^0.3 0.017 0.188 0.003 0.003 2.72 0.009 $77,271

Total Wear Material 0.591

$5,319,736


SGS FORM No. A263a-2/14/13 !


The fresh water cost estimate, shown in Table 21.18, is based on the fresh water pumps power consumption from overall water balance for the Redhawk Project, and the delivered water price of US$ 0.14 per ton will be used based on power consumption.

The repair materials and operating supplies is estimated using empirical factor based on total equipment installed cost, SGS recommends to use 4.0 percent for this 25,000 ton per day plant.

21.4 Tailings Storage Facility Capital Costs

Table 21.20 includes an estimate of capital costs for the proposed TSF. Components considered in the cost estimate include:

▪ Site preparation and topsoil salvage; ▪ TSF earthwork; ▪ Seepage collection pond construction; ▪ Underdrain system construction; ▪ Water Reclaim Works

Earthworks costs assume that the embankment will be constructed by the downstream raise method using locally borrowed fill. Unit rates are based on similar

Table 21.18 Water Cost Estimate

Usage

Fresh Water Pumps kWh per ton 1.80

Power Cost, $ per kWh $0.077

Fresh Water Pump Power Cost, $ per ton $0.14

Water Cost, $ per ton ore $0.14

Table 21.19 Repair Materials and Operating Supplies

Tonnage 25,000 tpd Source of Information

Total Equipment Installed Cost 53,470,806 Capital Cost Estimate

Maintenance Percentage 4.00% Other projects

Annual Maintenance Cost 2,138,832 Calculated

Cost per ton, US$ per ton ore 0.24 Calculated


SGS FORM No. A263a-2/14/13 !


activities at mining projects in the southwest US. A cost of $4.00/cubic yard (cy) is assumed for borrowing and placing local alluvial materials. A cost of $14.00/cy is assumed for materials that must be crushed and/or screened (liner bedding fill, drainage fill, and transitional fill materials).

To date, base metal tailings impoundments in Arizona have been permitted without geomembrane liners. The Copper Creek TSF is assumed to be un-lined. Construction costs include an allowance for a geomembrane liner to be placed on the upstream face of the embankment for control of the phreatic surface within the tailings dam. The seepage collection pond is assumed to meet Arizona Department of Environmental Quality (ADEQ) prescriptive design criteria for a process solution pond. Liner material and installation costs are assumed to be $0.80 and $0.60 per square foot for 80 and 60 mil geomembranes, respectively, based on typical installations.

Construction is assumed to occur in 5 campaigns occurring in Years minus 1, 1, 2, 4, and 8. Cost accuracy is on the order of plus or minus 40 percent.


SGS FORM No. A263a-2/14/13 !


Table 21.20 Tailings Storage Facility Capital Costs

!


SGS FORM No. A263a-2/14/13 !


Water reclaim system costs are based on recent vendor quotes for barge and pump systems of similar size.

Capital TSF costs do not include engineering, procurement and construction management (EPCM) costs. It is assumed that EPCM costs will be applied globally to the overall project cost estimate.

21.5 Diversion Tunnel Costs

Diversion tunnel costs are based on the assumption that a tunnel boring machine (TBM) will be used. Budgetary quotes for TBM equipment mobilization and set-up, tunneling and tunnel lining were provided by Frontier Kemper. Costs for inlet and outlet works are based on the requirements for typical, reinforced concrete headwall and wing wall structures and were developed with construction industry cost data (RS Means, Heavy Construction Cost Data, 26th Edition).

Table 21.21 contains the diversion tunnel cost estimate. EPCM costs are not included. Tunnel construction will not be required at project startup. It is estimated that subsidence crack propagation will not reach Copper Creek for approximately 6 years.


SGS FORM No. A263a-2/14/13 !


Table 21.21 Diversion Tunnel Cost Estimate

!


SGS FORM No. A263a-2/14/13 !


21.6 Reclamation and Closure

A closure and reclamation cost estimate for the TSF is included in Table 21.22. Costs assume a 5-year pre-closure monitoring period while the tailings in the interior of the TSF consolidate and surface traffic ability conditions improve. During this period, it is assumed that seepage will be circulated from the seepage collection pond to the TSF surface to enhance evaporation. The pre-closure costs include allowances for local monitoring well sampling and testing, and periodic maintenance of erosion controls by an earthworks contractor.

Reclamation is assumed to involve the placement of 24 inches of soil cover recovered from topsoil stockpiles. Due to potentially difficult conditions on the TSF surface, a cost of $5.00/cy is assumed for cover placement. The TSF reclamation cover and the downstream embankment slope will be reseeded. A permanent spillway will be required. Disposal operations can be modified during the final years of operation to push the free water pond to the vicinity of the designated spillway area.


SGS FORM No. A263a-2/14/13 !


Table 21.22 Tailings Storage Facility Closure & Reclamation Costs

!


SGS FORM No. A263a-2/14/13 !


22.0 ECONOMIC ANALYSIS

22.1 Metal Price Assumptions

Metal pricing for the financial model is summarized in Table 22.0.

22.2 Royalty

A sliding net returns royalty is payable to BHP on all production. Expenses after the product leaves the property are deducted from the gross value received. The sliding royalty is based upon the COMEX copper price as follows:

Less than US$ 0.80 per lb no royalty US$ 0.80 to US$ 0.99 per lb 1 percent royalty US$ 0.99 to US$ 1.10 per lb 2 percent royalty US$ 1.10 to US$ 1.20 per lb 2.5 percent royalty More than US$ 1.20 per lb 3 percent royalty

No advanced royalty payments are due.

22.3 Taxes

The cash flow analysis is pre-tax.

22.4 Financing

The economic analysis has been run on a basis of 100 percent equity financing.

22.5 Inflation

The economic analysis has been run with no inflation (constant January 2013 United States Dollars).

22.6 Economic Results

The economic analysis is summarized in the following Figures 22.1 and 22.2 and shown in Tables 22.1 to 22.13.

Table 22.0 Base Case Economic Model Metal Price Assumptions

US$

Copper, $ per lb. 3.00

Molybdenum, $ per lb MoO3 12.00

Silver, $ per oz 20.00


SGS FORM No. A263a-2/14/13 !


The financial analysis is pre-tax and based on a copper price of US$ 3.00 per pound. The pre-tax economic analysis results indicate that an IRR of 11.8 percent is achieved. The corresponding pre-tax NPV is US$ 457.0 million at a 5 percent discount rate, US$ 231.4 million at a 7.5 percent discount rate and US$ 79.3 million at a 10 percent discount rate. Payback of capital invested is achieved after 6.2 years of operation.

At a copper price of US$ 2.37 per pound a NPV of zero is produced at a zero percent discount rate (break even copper price). With a discount rate of 7.5 percent, a NPV of zero is produced at a copper price of US$ 2.73 per pound.

Sensitivity analysis indicating the effects of varying the capital and operating costs, recoveries, and metal prices are shown in Figures 22.1 and 22.2. Conventional financial criteria including the internal rate of return (IRR) and the net present value (NPV) at various discount rates are also indicated. As shown in Figures 22.1 and 22.2, the project is most sensitive to copper price and operating costs.


SGS FORM No. A263a-2/14/13 !


22.1 IRR SENSITIVITY (at US$ 3.00 per lb. Cu)

!

Sensitivity

Base Case -20% -15% -10% -5% Base 5% 10% 15% 20%

Copper Price

$3.00/lb 0.9% 4.2% 7.0% 9.5% 11.8% 14.0% 16.0% 17.9% 19.7%

Moly Price

$12.00/lb 11.3% 11.4% 11.6% 11.7% 11.8% 12.0% 12.1% 12.2% 12.4%

Copper Grade

0.948% 1.9% 4.8% 7.4% 9.7% 11.8% 13.8% 15.7% 17.5% 19.2%

Copper Recovery

92.00% 1.9% 4.8% 7.4% 9.7% 11.8% 13.8% 15.7%

Moly Recovery

78.00% 11.3% 11.4% 11.6% 11.7% 11.8% 12.0% 12.1% 12.2% 12.3%

Operating Cost 16.7% 15.5% 14.3% 13.1% 11.8% 10.5% 9.1% 7.6% 5.9%


SGS FORM No. A263a-2/14/13 !


FIGURE 22.2 NPV SENSITIVITY (at US$ 3.00 per lb. Cu)

!

Capital Cost 16.2% 15.0% 13.8% 12.8% 11.8% 10.9% 10.1% 9.3% 8.6%

Sensitivity

Base Case -20% -15% -10% -5% Base 5% 10% 15% 20%

Copper Price

$3.00/lb

($280,859.8)

($152,799.1)

($24,738.4)

$103,322.4

$231,383.1

$359,443.8

$487,504.6

$615,565.3

$743,626.0

Moly Price

$12.00/lb

$200,448.2

$208,181.9

$215,915.6

$223,649.4

$231,383.1

$239,116.8

$246,850.6

$254,584.3

$262,318.0

Copper Grade

0.948%

($241,407.4)

($123,209.8)

($5,012.2)

$113,185.5

$231,383.1

$349,580.7

$467,778.4

$585,976.0

$704,173.6

Copper Recovery

92.00%

($241,407.4)

($123,209.8)

($5,012.2)

$113,185.5

$231,383.1

$349,580.7

$467,778.4

Moly Recovery

78.00%

$201,479.3

$208,955.3

$216,431.2

$223,907.2

$231,383.1

$238,859.0

$246,335.0

$253,810.9

$261,286.9

Operating Cost

$535,778.4

$459,679.6

$383,580.8

$307,481.9

$231,383.1

$155,284.3

$79,185.4

$3,086.6

($73,012.2)


SGS FORM No. A263a-2/14/13 !


!

!

Capital Cost

$396,804.0

$355,448.7

$314,093.5

$272,738.3

$231,383.1

$190,027.9

$148,672.7

$107,317.5

$65,962

Table 22-1 Mine Production Schedule

YEAR -4 -3 -2 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 TotalORE PRODUCTION

Underground productionOre tons mined (000's) 7,200 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 4,690 155,890

Copper grade, % total copper 0.945 0.889 0.915 0.741 0.726 0.726 0.733 0.728 0.727 0.712 0.754 0.754 0.754 0.754 0.730 0.769 0.785 0.726 0.770Silver grade, opt 0.09 0.10 0.08 0.06 0.06 0.06 0.07 0.07 0.07 0.07 0.03 0.03 0.03 0.03 0.07 0.05 0.03 0.02 0.056Molybdenum grade, % 0.018 0.019 0.020 0.015 0.014 0.014 0.014 0.014 0.015 0.023 0.012 0.012 0.012 0.012 0.021 0.017 0.011 0.014 0.015

Copper contained, lbs. (000's) 136,080 160,020 164,700 133,380 130,680 130,680 131,940 131,040 130,860 128,160 135,720 135,720 135,720 135,720 131,400 138,420 141,300 68,099 2,399,639Silver contained, ounces 669,600 864,000 711,000 495,000 495,000 495,000 585,000 630,000 648,000 603,000 306,000 306,000 306,000 306,000 612,000 405,000 252,000 103,180 8,791,780Molybdenum, contained, lbs. (000's) 2,592 3,420 3,600 2,700 2,520 2,520 2,520 2,520 2,700 4,140 2,160 2,160 2,160 2,160 3,780 3,060 1,980 1,313 48,005

Table 22-2 Ore Milling Schedule

YEAR -4 -3 -2 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 Total

Ore to be milledTons of ore mined (000's) 7,200 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 4,690 155,890Transportation losses,% 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0Tons of ore mined (000's) 7,200 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 4,690 155,890

Copper grade, % 0.945 0.889 0.915 0.741 0.726 0.726 0.733 0.728 0.727 0.712 0.754 0.754 0.754 0.754 0.730 0.769 0.785 0.726 0.770Silver grade, opt 0.09 0.10 0.08 0.06 0.06 0.06 0.07 0.07 0.07 0.07 0.03 0.03 0.03 0.03 0.07 0.05 0.03 0.02 0.056Molybdenum grade, % 0.018 0.019 0.020 0.015 0.014 0.014 0.014 0.014 0.015 0.023 0.012 0.012 0.012 0.012 0.021 0.017 0.011 0.014 0.015


Table 22-3 Flotation Concentrator Feed Production Schedule

YEAR -4 -3 -2 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 Total

FlotationOre ProcessedTons of ore processed (000's) 7,200 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 4,690 155,890Copper grade, % 0.945 0.889 0.915 0.741 0.726 0.726 0.733 0.728 0.727 0.712 0.754 0.754 0.754 0.754 0.730 0.769 0.785 0.726 0.770Gold grade, opt 0.000 0.000 0.000 0.000 0.000 0.000 0.000 0.000 0.000 0.000 0.000 0.000 0.000 0.000 0.000 0.000 0.000 0.000 0.000Silver grade, opt 0.093 0.096 0.079 0.055 0.055 0.055 0.065 0.070 0.072 0.067 0.034 0.034 0.034 0.034 0.068 0.045 0.028 0.022 0.056Molybdenum grade, % 0.018 0.019 0.020 0.015 0.014 0.014 0.014 0.014 0.015 0.023 0.012 0.012 0.012 0.012 0.021 0.017 0.011 0.014 0.015

Total Material ProcessedTons of ore processed (000's) 7,200 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 4,690 155,890Copper grade, % 0.945 0.889 0.915 0.741 0.726 0.726 0.733 0.728 0.727 0.712 0.754 0.754 0.754 0.754 0.730 0.769 0.785 0.726 0.770Gold grade, opt 0.000 0.000 0.000 0.000 0.000 0.000 0.000 0.000 0.000 0.000 0.000 0.000 0.000 0.000 0.000 0.000 0.000 0.000 0.000Silver grade, opt 0.093 0.096 0.079 0.055 0.055 0.055 0.065 0.070 0.072 0.067 0.034 0.034 0.034 0.034 0.068 0.045 0.028 0.022 0.056Molybdenum grade, % 0.018 0.019 0.020 0.015 0.014 0.014 0.014 0.014 0.015 0.023 0.012 0.012 0.012 0.012 0.021 0.017 0.011 0.014 0.015

Copper contained, lbs. (000's) 136,080 160,020 164,700 133,380 130,680 130,680 131,940 131,040 130,860 128,160 135,720 135,720 135,720 135,720 131,400 138,420 141,300 68,099 2,399,639Gold contained, ounces 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0Silver contained, ounces 669,600 864,000 711,000 495,000 495,000 495,000 585,000 630,000 648,000 603,000 306,000 306,000 306,000 306,000 612,000 405,000 252,000 103,180 8,791,780Molybdenum, contained, lbs. (000's) 2,592 3,420 3,600 2,700 2,520 2,520 2,520 2,520 2,700 4,140 2,160 2,160 2,160 2,160 3,780 3,060 1,980 1,313 48,005

Copper Recovery, % 92.00 92.00 92.00 92.00 92.00 92.00 92.00 92.00 92.00 92.00 92.00 92.00 92.00 92.00 92.00 92.00 92.00 92.00 92.00Silver Recovery, % 50.00 50.00 50.00 50.00 50.00 50.00 50.00 50.00 50.00 50.00 50.00 50.00 50.00 50.00 50.00 50.00 50.00 50.00 50.00Molybdenum Recovery, % 78.00 78.00 78.00 78.00 78.00 78.00 78.00 78.00 78.00 78.00 78.00 78.00 78.00 78.00 78.00 78.00 78.00 78.00 78.00

Table 22-4 Flotation Concentrate Production Schedule

YEAR -4 -3 -2 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 TotalFLOTATION CONCENTRATE PRODUCED

Flotation concentrate produced -Tons of Cu. Concentrate produced 208,656 245,364 252,540 204,516 200,376 200,376 202,308 200,928 200,652 196,512 208,104 208,104 208,104 208,104 201,480 212,244 216,660 104,418 3,679,446Copper grade, % 30.00 30.00 30.00 30.00 30.00 30.00 30.00 30.00 30.00 30.00 30.00 30.00 30.00 30.00 30.00 30.00 30.00 30.00 30.000Silver grade, opt 1.605 1.761 1.408 1.210 1.235 1.235 1.446 1.568 1.615 1.534 0.735 0.735 0.735 0.735 1.519 0.954 0.582 0.494 1.195

Copper contained, lbs. (000's) 125,194 147,218 151,524 122,710 120,226 120,226 121,385 120,557 120,391 117,907 124,862 124,862 124,862 124,862 120,888 127,346 129,996 62,651 2,207,668Silver contained, ounces 334,800 432,000 355,500 247,500 247,500 247,500 292,500 315,000 324,000 301,500 153,000 153,000 153,000 153,000 306,000 202,500 126,000 51,590 4,395,890

Tons of Molybdenum concentrate produced 2,151 2,838 2,987 2,240 2,091 2,091 2,091 2,091 2,240 3,435 1,792 1,792 1,792 1,792 3,137 2,539 1,643 1,090 39,834Molybdenum grade, % 47.00 47.00 47.00 47.00 47.00 47.00 47.00 47.00 47.00 47.00 47.00 47.00 47.00 47.00 47.00 47.00 47.00 47.00 47.00Molybdenum, contained, lbs. (000's) 2,022 2,668 2,808 2,106 1,966 1,966 1,966 1,966 2,106 3,229 1,685 1,685 1,685 1,685 2,948 2,387 1,544 1,024 37,444

Table 22-5 Concentrates Smelted Schedule

YEAR -4 -3 -2 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 TotalFLOTATION CONCENTRATE SMELTED

Transportation losses, % 0.25 0.25 0.25 0.25 0.25 0.25 0.25 0.25 0.25 0.25 0.25 0.25 0.25 0.25 0.25 0.25 0.25 0.25

Flotation concentrate smelted -Tons of Cu. concentrate smelted 208,134 244,751 251,909 204,005 199,875 199,875 201,802 200,426 200,150 196,021 207,584 207,584 207,584 207,584 200,976 211,713 216,118 104,157 3,670,248Copper grade, % 30.00 30.00 30.00 30.00 30.00 30.00 30.00 30.00 30.00 30.00 30.00 30.00 30.00 30.00 30.00 30.00 30.00 30.00 30.000Silver grade, opt 1.605 1.761 1.408 1.210 1.235 1.235 1.446 1.568 1.615 1.534 0.735 0.735 0.735 0.735 1.519 0.954 0.582 0.494 1.195Molybdenum grade, % 47.00 47.00 47.00 47.00 47.00 47.00 47.00 47.00 47.00 47.00 47.00 47.00 47.00 47.00 47.00 47.00 47.00 47.00


Table 22-6 Payable Metal and Commodity Schedule

YEAR -4 -3 -2 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 TotalPAYABLE, %

Flotation concentrate (Smelter payable) Copper, % 96.67 96.67 96.67 96.67 96.67 96.67 96.67 96.67 96.67 96.67 96.67 96.67 96.67 96.67 96.67 96.67 96.67 96.67 96.67 Silver, % 35.79 41.04 27.51 16.50 18.09 18.09 29.29 34.40 36.17 33.08 21.30 Moly, % 85.17 85.17 85.17 85.17 85.17 85.17 85.17 85.17 85.17 85.17 85.17 85.17 85.17 85.17 85.17 85.17 85.17 85.17 85.17

PAYABLE METAL

Flotation concentrates Copper, lbs. (000's) 120,718 141,955 146,107 118,323 115,928 115,928 117,045 116,247 116,087 113,692 120,399 120,399 120,399 120,399 116,566 122,794 125,349 60,411 2,128,744 Silver, ounces 119,537 176,861 97,567 40,733 44,656 44,656 85,468 108,097 116,888 99,489 0 0 0 0 0 0 0 0 933,953 Molybdenum, lbs. (000's) 1,722 2,272 2,392 1,794 1,674 1,674 1,674 1,674 1,794 2,750 1,435 1,435 1,435 1,435 2,511 2,033 1,315 872 31,890

COMMODITY PRICES Copper ($ per lb.) 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 Silver ($ per oz) 20.00 20.00 20.00 20.00 20.00 20.00 20.00 20.00 20.00 20.00 20.00 20.00 20.00 20.00 20.00 20.00 20.00 20.00 20.00 Molybdenum ($ per lb. in MoO3) 12.00 12.00 12.00 12.00 12.00 12.00 12.00 12.00 12.00 12.00 12.00 12.00 12.00 12.00 12.00 12.00 12.00 12.00 12.00


SGS FORM No. A263a-2/14/13 !


!

!

!

Table 22-7 Net Smelter Return Schedule

YEAR -4 -3 -2 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 Total

NET SMELTER RETURNGross revenue ($ 000's): Copper 362,154 425,866 438,321 354,968 347,783 347,783 351,136 348,741 348,262 341,076 361,196 361,196 361,196 361,196 349,699 368,381 376,046 181,233 6,386,231 Silver 2,391 3,537 1,951 815 893 893 1,709 2,162 2,338 1,990 0 0 0 0 0 0 0 0 18,679 Molybdenum 20,663 27,263 28,698 21,524 20,089 20,089 20,089 20,089 21,524 33,003 17,219 17,219 17,219 17,219 30,133 24,394 15,784 10,468 382,685 Total 385,207 456,667 468,971 377,307 368,765 368,765 372,934 370,991 372,123 376,069 378,415 378,415 378,415 378,415 379,832 392,775 391,830 191,702 6,787,595

Treatment charges (rates per unit)

Flotation Concentrates Smelting ($/ton of concentrate) 60 60 60 60 60 60 60 60 60 60 60 60 60 60 60 60 60 60 60 Refining-Cu ($/Cu. lb.) 0.06 0.06 0.06 0.06 0.06 0.06 0.06 0.06 0.06 0.06 0.06 0.06 0.06 0.06 0.06 0.06 0.06 0.06 0.06 Copper Price Participation ($/ lb. Cu. ) 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 #DIV/0! Ag Refining-Cu Conc.($/Ag. oz) 0.50 0.50 0.50 0.50 0.50 0.50 0.50 0.50 0.50 0.50 0.50 0.50 0.50 0.50 0.50 0.50 0.50 0.50 0.50 Moly Roast & Leach, $ per pound 0.40 0.40 0.40 0.40 0.40 0.40 0.40 0.40 0.40 0.40 0.40 0.40 0.40 0.40 0.40 0.40 0.40 0.40 0.40

Treatment charges ($ 000's)

Flotation Concentrates Smelting 12,488 14,685 15,115 12,240 11,993 11,993 12,108 12,026 12,009 11,761 12,455 12,455 12,455 12,455 12,059 12,703 12,967 6,249 220,215 Copper Refining 7,243 8,517 8,766 7,099 6,956 6,956 7,023 6,975 6,965 6,822 7,224 7,224 7,224 7,224 6,994 7,368 7,521 3,625 127,725 Ag Refining-Cu Conc. 60 88 49 20 22 22 43 54 58 50 0 0 0 0 0 0 0 0 467 Moly Roast & Leach 689 909 957 717 670 670 670 670 717 1,100 574 574 574 574 1,004 813 526 349 12,756 Total 20,480 24,200 24,886 20,077 19,640 19,640 19,843 19,724 19,750 19,733 20,253 20,253 20,253 20,253 20,057 20,884 21,014 10,223 361,163

Net Smelter Return ($ 000's) 364,728 432,467 444,084 357,229 349,124 349,124 353,091 351,267 352,373 356,336 358,162 358,162 358,162 358,162 359,775 371,891 370,816 181,479 6,426,432

Table 22-8 Cash Operating Cost Schedule

YEAR -4 -3 -2 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 Total

COSTS ($ 000's)

OPERATING COST

Underground mining 143,208 171,450 164,250 157,050 157,050 157,050 157,050 157,050 157,050 157,050 157,050 157,050 157,050 157,050 157,050 157,050 157,050 81,841 2,759,449

Milling and ore handling 35,837 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 23,344 775,926

Freight to mill/smelter 5,216 6,134 6,314 5,113 5,009 5,009 5,058 5,023 5,016 4,913 5,203 5,203 5,203 5,203 5,037 5,306 5,417 2,610 91,986

General & administrative 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 96,228TOTAL OPERATING COSTS 189,608 227,727 220,706 212,305 212,202 212,202 212,250 212,216 212,209 212,105 212,395 212,395 212,395 212,395 212,230 212,499 212,609 113,141 3,723,589

TREATMENT CHARGES Smelting 12,488 14,685 15,115 12,240 11,993 11,993 12,108 12,026 12,009 11,761 12,455 12,455 12,455 12,455 12,059 12,703 12,967 6,249 220,215 Refining 7,243 8,517 8,766 7,099 6,956 6,956 7,023 6,975 6,965 6,822 7,224 7,224 7,224 7,224 6,994 7,368 7,521 3,625 127,725 Ag Refining-Cu Conc. 60 88 49 20 22 22 43 54 58 50 0 0 0 0 0 0 0 0 467 Moly Roast & Leach, $ per pound 689 909 957 717 670 670 670 670 717 1,100 574 574 574 574 1,004 813 526 349 12,756 Total treatment charges 20,480 24,200 24,886 20,077 19,640 19,640 19,843 19,724 19,750 19,733 20,253 20,253 20,253 20,253 20,057 20,884 21,014 10,223 361,163

TOTAL CASH OPERATING COST 210,087 251,926 245,592 232,383 231,842 231,842 232,093 231,940 231,959 231,838 232,648 232,648 232,648 232,648 232,287 233,382 233,623 123,364 4,084,751

BY-PRODUCT CREDITS 22,305 29,803 29,644 21,601 20,290 20,290 21,086 21,527 23,086 33,843 16,645 16,645 16,645 16,645 29,129 23,580 15,258 10,120 388,141

TOTAL CASH OPERATING COST, NET OF BY-PRODUCT CREDIT ($ 000's) 187,782 222,123 215,948 210,782 211,552 211,552 211,008 210,413 208,873 197,995 216,003 216,003 216,003 216,003 203,158 209,802 218,365 113,244 3,696,610Copper, lbs. Sold (000's) From concentrate 120,718 141,955 146,107 118,323 115,928 115,928 117,045 116,247 116,087 113,692 120,399 120,399 120,399 120,399 116,566 122,794 125,349 60,411 2,128,744

UNIT COST ($/Cu. lb.)OPERATING COST Underground Mining 1.1863 1.2078 1.1242 1.3273 1.3547 1.3547 1.3418 1.3510 1.3529 1.3814 1.3044 1.3044 1.3044 1.3044 1.3473 1.2790 1.2529 1.3547 1.2963 Milling and ore handling 0.2969 0.3156 0.3066 0.3786 0.3864 0.3864 0.3827 0.3854 0.3859 0.3940 0.3721 0.3721 0.3721 0.3721 0.3843 0.3648 0.3574 0.3864 0.3645 Freight to mill/smelter 0.0432 0.0432 0.0432 0.0432 0.0432 0.0432 0.0432 0.0432 0.0432 0.0432 0.0432 0.0432 0.0432 0.0432 0.0432 0.0432 0.0432 0.0432 0.0432 General & administrative 0.0443 0.0377 0.0366 0.0452 0.0461 0.0461 0.0457 0.0460 0.0461 0.0470 0.0444 0.0444 0.0444 0.0444 0.0459 0.0435 0.0426 0.0885 0.0452 Total Operating cost 1.5707 1.6042 1.5106 1.7943 1.8305 1.8305 1.8134 1.8256 1.8280 1.8656 1.7641 1.7641 1.7641 1.7641 1.8207 1.7305 1.6961 1.8728 1.7492

TREATMENT CHARGES Smelting 0.1034 0.1034 0.1034 0.1034 0.1034 0.1034 0.1034 0.1034 0.1034 0.1034 0.1034 0.1034 0.1034 0.1034 0.1034 0.1034 0.1034 0.1034 0.1034 Refining 0.0600 0.0600 0.0600 0.0600 0.0600 0.0600 0.0600 0.0600 0.0600 0.0600 0.0600 0.0600 0.0600 0.0600 0.0600 0.0600 0.0600 0.0600 0.0600 Ag Refining-Cu Conc. 0.0005 0.0006 0.0003 0.0002 0.0002 0.0002 0.0004 0.0005 0.0005 0.0004 0.0000 0.0000 0.0000 0.0000 0.0000 0.0000 0.0000 0.0000 0.0002 Moly Roast & Leach, $ per pound 0.0057 0.0064 0.0065 0.0061 0.0058 0.0058 0.0057 0.0058 0.0062 0.0097 0.0048 0.0048 0.0048 0.0048 0.0086 0.0066 0.0042 0.0058 0.0060 Total treatment charges 0.1696 0.1705 0.1703 0.1697 0.1694 0.1694 0.1695 0.1697 0.1701 0.1736 0.1682 0.1682 0.1682 0.1682 0.1721 0.1701 0.1676 0.1692 0.1697

TOTAL CASH OPERATING COST 1.7403 1.7747 1.6809 1.9640 1.9999 1.9999 1.9829 1.9952 1.9981 2.0392 1.9323 1.9323 1.9323 1.9323 1.9927 1.9006 1.8638 2.0421 1.9189

BY-PRODUCT CREDITS 0.1848 0.2099 0.2029 0.1826 0.1750 0.1750 0.1802 0.1852 0.1989 0.2977 0.1382 0.1382 0.1382 0.1382 0.2499 0.1920 0.1217 0.1675 0.1823

TOTAL CASH OPERATING COST, NET OF BY-PRODUCT CREDIT ($/Cu. Lb.) 1.5555 1.5647 1.4780 1.7814 1.8249 1.8249 1.8028 1.8101 1.7993 1.7415 1.7941 1.7941 1.7941 1.7941 1.7429 1.7086 1.7421 1.8746 1.7365

Table 22-9 Cash Operating Cost Per Ton Mined Schedule

YEAR -4 -3 -2 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 Total

Total tons milled (000's) 7,200 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 4,690 155,890

UNIT COSTS ($/ton milled)OPERATING COST Underground Mining 19.89 19.05 18.25 17.45 17.45 17.45 17.45 17.45 17.45 17.45 17.45 17.45 17.45 17.45 17.45 17.45 17.45 17.45 17.70 Milling and ore handling 4.98 4.98 4.98 4.98 4.98 4.98 4.98 4.98 4.98 4.98 4.98 4.98 4.98 4.98 4.98 4.98 4.98 4.98 4.98 Freight to mill/smelter 0.72 0.68 0.70 0.57 0.56 0.56 0.56 0.56 0.56 0.55 0.58 0.58 0.58 0.58 0.56 0.59 0.60 0.56 0.59 General & administrative 0.74 0.59 0.59 0.59 0.59 0.59 0.59 0.59 0.59 0.59 0.59 0.59 0.59 0.59 0.59 0.59 0.59 1.14 0.62 Total Operating cost 26.33 25.30 24.52 23.59 23.58 23.58 23.58 23.58 23.58 23.57 23.60 23.60 23.60 23.60 23.58 23.61 23.62 24.12 23.89

TREATMENT CHARGES Smelting 1.73 1.63 1.68 1.36 1.33 1.33 1.35 1.34 1.33 1.31 1.38 1.38 1.38 1.38 1.34 1.41 1.44 1.33 1.41 Refining 1.01 0.95 0.97 0.79 0.77 0.77 0.78 0.77 0.77 0.76 0.80 0.80 0.80 0.80 0.78 0.82 0.84 0.77 0.82 Ag Refining-Cu Conc. 0.01 0.01 0.01 0.00 0.00 0.00 0.00 0.01 0.01 0.01 - - - - - - - - 0.00 Moly Roast & Leach, $ per pound 0.10 0.10 0.11 0.08 0.07 0.07 0.07 0.07 0.08 0.12 0.06 0.06 0.06 0.06 0.11 0.09 0.06 0.07 0.08 Total treatment charges 2.84 2.69 2.77 2.23 2.18 2.18 2.20 2.19 2.19 2.19 2.25 2.25 2.25 2.25 2.23 2.32 2.33 2.18 2.32

TOTAL CASH OPERATING COST 29.18 27.99 27.29 25.82 25.76 25.76 25.79 25.77 25.77 25.76 25.85 25.85 25.85 25.85 25.81 25.93 25.96 26.30 26.20

BY-PRODUCT CREDITS 3.10 3.31 3.29 2.40 2.25 2.25 2.34 2.39 2.57 3.76 1.85 1.85 1.85 1.85 3.24 2.62 1.70 2.16 2.49

TOTAL CASH OPERATING COST, NET OF BY-PRODUCT CREDIT ($/ton milled) 26.08 24.68 23.99 23.42 23.51 23.51 23.45 23.38 23.21 22.00 24.00 24.00 24.00 24.00 22.57 23.31 24.26 24.15 23.71

Table 22-10 Byproduct Credit Schedule

YEAR -4 -3 -2 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 Total

Gross revenue ($ 000's): Molybdenum 20,663 27,263 28,698 21,524 20,089 20,089 20,089 20,089 21,524 33,003 17,219 17,219 17,219 17,219 30,133 24,394 15,784 10,468 382,685 Silver 2,391 3,537 1,951 815 893 893 1,709 2,162 2,338 1,990 0 0 0 0 0 0 0 0 18,679 Total 23,054 30,801 30,650 22,338 20,982 20,982 21,798 22,251 23,861 34,993 17,219 17,219 17,219 17,219 30,133 24,394 15,784 10,468 401,364

By-product treatment charges ($ 000's):

Moly Concentrate Processing Moly Roast & Leach @ $ 0.4 per pound 689 909 957 717 670 670 670 670 717 1100 574 574 574 574 1004 813 526 349 12,756

Copper concentrate by-products Refining-Ag 60 88 49 20 22 22 43 54 58 50 0 0 0 0 0 0 0 0 467

Total by-product treatment charges 749 997 1,005 738 692 692 712 724 776 1,150 574 574 574 574 1,004 813 526 349 13,223

Net By-Product Credit 22,305 29,803 29,644 21,601 20,290 20,290 21,086 21,527 23,086 33,843 16,645 16,645 16,645 16,645 29,129 23,580 15,258 10,120 388,141

Net By-Product Credit ($/Cu. lb.) 0.1848 0.2099 0.2029 0.1826 0.1750 0.1750 0.1802 0.1852 0.1989 0.2977 0.1382 0.1382 0.1382 0.1382 0.2499 0.1920 0.1217 0.1675 0.1823


SGS FORM No. A263a-2/14/13 !


!

!

Table 22-11 Cash Operating Calculation Schedule

YEAR -4 -3 -2 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 Total

MINING COSTS

Total MinedOre tons mined (000's) 7,200 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 4,690 155,890

Mining cost ($ 000's) 143,208 171,450 164,250 157,050 157,050 157,050 157,050 157,050 157,050 157,050 157,050 157,050 157,050 157,050 157,050 157,050 157,050 81,841 2,759,449Mining cost ($/ton of ore) 19.89 19.05 18.25 17.45 17.45 17.45 17.45 17.45 17.45 17.45 17.45 17.45 17.45 17.45 17.45 17.45 17.45 17.45 17.70

PROCESSING COSTSMillingTons milled (000's) 7,200 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 4,690 155,890Nominal daily mill tonnage 20,000 25,000 25,000 25,000 25,000 25,000 25,000 25,000 25,000 25,000 25,000 25,000 25,000 25,000 25,000 25,000 25,000 13,028Processing cost ($ 000's) 35,837 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 23,344 775,926Operating cost ($/ton milled) 4.98 4.98 4.98 4.98 4.98 4.98 4.98 4.98 4.98 4.98 4.98 4.98 4.98 4.98 4.98 4.98 4.98 4.98 4.98

Freight to smelter cost ($000's) 5,216 6,134 6,314 5,113 5,009 5,009 5,058 5,023 5,016 4,913 5,203 5,203 5,203 5,203 5,037 5,306 5,417 2,610 91,986Freight to smelter cost ($/ton of conc.) 25.00 25.00 25.00 25.00 25.00 25.00 25.00 25.00 25.00 25.00 25.00 25.00 25.00 25.00 25.00 25.00 25.00 25.00 25.00

General and AdministrativeOnsite cost ($/ ore ton mined) 0.74 0.59 0.59 0.59 0.59 0.59 0.59 0.59 0.59 0.59 0.59 0.59 0.59 0.59 0.59 0.59 0.59 1.14 0.62

Onsite cost ($ 000's) 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 96,228Land Payments ($ 000's) 0Total ($ 000's) 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 96,228

Total costs Mining ($ 000's) 143,208 171,450 164,250 157,050 157,050 157,050 157,050 157,050 157,050 157,050 157,050 157,050 157,050 157,050 157,050 157,050 157,050 81,841 2,759,449Paste production ($ 000's) 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0Ore processing ($ 000's) 35,837 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 23,344 775,926Freight to mill/smelter ($ 000's) 5,216 6,134 6,314 5,113 5,009 5,009 5,058 5,023 5,016 4,913 5,203 5,203 5,203 5,203 5,037 5,306 5,417 2,610 91,986General & administrative ($ 000's) 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 96,228Total costs ($ 000's) 189,608 227,727 220,706 212,305 212,202 212,202 212,250 212,216 212,209 212,105 212,395 212,395 212,395 212,395 212,230 212,499 212,609 113,141 3,723,589

BHP royalty rate, percent 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00BHP royalty ($ 000's) 10,942 12,974 13,323 10,717 10,474 10,474 10,593 10,538 10,571 10,690 10,745 10,745 10,745 10,745 10,793 11,157 11,124 5,444 192,793

Total royalty ($ 000's) 10,942 12,974 13,323 10,717 10,474 10,474 10,593 10,538 10,571 10,690 10,745 10,745 10,745 10,745 10,793 11,157 11,124 5,444 192,793

Table 22-12 Capital Cost Schedule

YEAR -4 -3 -2 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 Total

CAPITAL EXPENDITURE ($ 000's)

Mine development

Mine development cost total 75,397 60,806 146,909 258,805 5,720 5,720 5,720 5,720 5,720 40,820 5,720 5,720 5,720 40,820 5,720 5,720 5,720 39,247 5,720 5,720 5,720 0 742,885

Primary 12' Diversion Tunnel 37,906 37,906

Tailing Dam with Contingency 16,428 8,467 15,739 22,295 41,047 103,975

Closure Costs 16,411 16,411

Plant and equipmentTotal Direct Cost 0 133,374 44,458 177,832Total Indirect Cost 10,861 21,722 10,861 43,443Contingency 0 38,723 38,723 77,446

Total Plant Cost 10,861 193,819 94,042 298,721

Total mine and mill capital 75,397 71,667 340,728 369,275 14,187 21,459 5,720 28,015 5,720 40,820 5,720 84,673 5,720 40,820 5,720 5,720 5,720 39,247 5,720 5,720 5,720 0 16,411 1,199,899

Table 22-13 Cash Flow Schedule

YEAR -4 -3 -2 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 Total

ORE PRODUCTION

Total mining productionOre tons mined (000's) 7,200 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 9,000 4,690 155,890Copper grade, % 0.945 0.889 0.915 0.741 0.726 0.726 0.733 0.728 0.727 0.712 0.754 0.754 0.754 0.754 0.730 0.769 0.785 0.726 0.770Silver grade, opt 0.093 0.096 0.079 0.055 0.055 0.055 0.065 0.070 0.072 0.067 0.034 0.034 0.034 0.034 0.068 0.045 0.028 0.022 0.056Molybdenum grade, % 0.018 0.019 0.020 0.015 0.014 0.014 0.014 0.014 0.015 0.023 0.012 0.012 0.012 0.012 0.021 0.017 0.011 0.014 0.015

PAYABLE METAL

TotalCopper, lbs. (000's) 120,718 141,955 146,107 118,323 115,928 115,928 117,045 116,247 116,087 113,692 120,399 120,399 120,399 120,399 116,566 122,794 125,349 60,411 2,128,744Silver, ounces 119,537 176,861 97,567 40,733 44,656 44,656 85,468 108,097 116,888 99,489 0 0 0 0 0 0 0 0 933,953Molybdenum, lbs. (000's) 1,722 2,272 2,392 1,794 1,674 1,674 1,674 1,674 1,794 2,750 1,435 1,435 1,435 1,435 2,511 2,033 1,315 872 31,890

COMMODITY PRICESCopper ($ per lb.) 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 3.00 Molybdenum ($ per lb. in MoO3) 12.00 12.00 12.00 12.00 12.00 12.00 12.00 12.00 12.00 12.00 12.00 12.00 12.00 12.00 12.00 12.00 12.00 12.00 12.00Silver ($ per oz) 20.00 20.00 20.00 20.00 20.00 20.00 20.00 20.00 20.00 20.00 20.00 20.00 20.00 20.00 20.00 20.00 20.00 20.00 20.00

NET SMELTER RETURN ($ 000's)Gross revenue Copper 362,154 425,866 438,321 354,968 347,783 347,783 351,136 348,741 348,262 341,076 361,196 361,196 361,196 361,196 349,699 368,381 376,046 181,233 6,386,231 Molybdenum 20,663 27,263 28,698 21,524 20,089 20,089 20,089 20,089 21,524 33,003 17,219 17,219 17,219 17,219 30,133 24,394 15,784 10,468 382,685 Silver 2,391 3,537 1,951 815 893 893 1,709 2,162 2,338 1,990 0 0 0 0 0 0 0 0 18,679 Total 385,207 456,667 468,971 377,307 368,765 368,765 372,934 370,991 372,123 376,069 378,415 378,415 378,415 378,415 379,832 392,775 391,830 191,702 6,787,595

Treatment charges ($ 000's)Total Smelting 12,488 14,685 15,115 12,240 11,993 11,993 12,108 12,026 12,009 11,761 12,455 12,455 12,455 12,455 12,059 12,703 12,967 6,249 220,215 Refining 7,243 8,517 8,766 7,099 6,956 6,956 7,023 6,975 6,965 6,822 7,224 7,224 7,224 7,224 6,994 7,368 7,521 3,625 127,725 Ag Refining-Cu Conc. 60 88 49 20 22 22 43 54 58 50 0 0 0 0 0 0 0 0 467 Moly Roast & Leach, $ per pound 689 909 957 717 670 670 670 670 717 1,100 574 574 574 574 1,004 813 526 349 12,756 Total treatment charges 20,480 24,200 24,886 20,077 19,640 19,640 19,843 19,724 19,750 19,733 20,253 20,253 20,253 20,253 20,057 20,884 21,014 10,223 361,163

Net Smelter Return 364,728 432,467 444,084 357,229 349,124 349,124 353,091 351,267 352,373 356,336 358,162 358,162 358,162 358,162 359,775 371,891 370,816 181,479 6,426,432

OPERATING COST ($ 000's) Underground mining 143,208 171,450 164,250 157,050 157,050 157,050 157,050 157,050 157,050 157,050 157,050 157,050 157,050 157,050 157,050 157,050 157,050 81,841 2,759,449 Milling and ore handling 35,837 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 44,797 23,344 775,926 Freight to mill/smelter 5,216 6,134 6,314 5,113 5,009 5,009 5,058 5,023 5,016 4,913 5,203 5,203 5,203 5,203 5,037 5,306 5,417 2,610 91,986 General & administrative 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 5,346 96,228 Total operating cost 189,608 227,727 220,706 212,305 212,202 212,202 212,250 212,216 212,209 212,105 212,395 212,395 212,395 212,395 212,230 212,499 212,609 113,141 3,723,589

CAPITAL COST ($ 000's)Mine development cost total 75,397 60,806 146,909 258,805 5,720 5,720 5,720 5,720 5,720 40,820 5,720 5,720 5,720 40,820 5,720 5,720 5,720 39,247 5,720 5,720 5,720 0 742,885Primary 12' Diversion Tunnel 37,906 37,906Tailing Dam with Contingency 0 0 0 16,428 8,467 15,739 0 22,295 0 0 0 41,047 0 0 0 0 0 0 0 0 0 0 103,975Closure Costs 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 16,411 16,411Plant and equipment 0 10,861 193,819 94,042 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 298,721Total Capital Costs 75,397 71,667 340,728 369,275 14,187 21,459 5,720 28,015 5,720 40,820 5,720 84,673 5,720 40,820 5,720 5,720 5,720 39,247 5,720 5,720 5,720 0 16,411 1,199,899

Royalties NSR and Fees 10,942 12,974 13,323 10,717 10,474 10,474 10,593 10,538 10,571 10,690 10,745 10,745 10,745 10,745 10,793 11,157 11,124 5,444 192,793

PRE- TAX CASH FLOW ($ 000's) -75,397 -71,667 -340,728 -369,275 149,991 170,308 204,336 106,192 120,729 85,629 124,528 43,840 123,873 92,721 129,302 129,302 129,302 95,775 131,032 142,516 141,362 62,894 -16,411 1,310,152Cumulative Cash Flow ($ 000's) -75,397 -147,064 -487,792 -857,067 -707,076 -536,768 -332,432 -226,240 -105,512 -19,883 104,645 148,485 272,358 365,079 494,381 623,682 752,984 848,759 979,791 1,122,307 1,263,669 1,326,563 1,310,152Payback, operating years 6.2 7.2 7.2Pre- Tax Net Present Value

0.0% $ x 1000 $1,310,1525.0% $ x 1000 $456,9947.5% $ x 1000 $231,38310.0% $ x 1000 $79,31115.0% $ x 1000 ($94,372)

Pre- Tax Internal Rate of ReturnCalculated IRR 11.8%


SGS FORM No. A263a-2/14/13 !

Redhawk Copper, Inc. - Copper Creek PEA 25,000 TPD Mill Page !24123.0 ADJACENT PROPERTIES

The Copper Creek Property lies within the porphyry copper metallogenic province of the southwestern United States. There are several large adjacent properties that have been or still are major producers of copper. The San Manuel/Kalamazoo deposit lies within 15 miles of the Copper Creek Property and the Ray Mine lies approximately 50 miles north of the property. San Manuel was a major underground copper mine until its closure, with production in excess of 50,000 short tons per day. ASARCO Ray is a large open pit copper mine that is still in operation.

Within the Copper Creek Property, there has been historic copper production as discussed in Section 6. Mining in the Childs-Aldwinkle breccia was as recently as 1965 with the majority of mining done between 1933 and 1938. This mining occurred in the central and southern fingers of the breccia pipe. The resource estimate for the Childs-Aldwinkle presented in this report is in the north finger and the breccia body below the three fingers. There is no evidence of historic mining in the area of the resource estimate. Mining in the Old Reliable occurred in 1972 when the upper portion of it was rubblized with one major blast and the broken rock was leached for copper. The leaching operation continued through 1981 with about 12 million pounds of copper recovered. Some of the post leaching drilling shows remaining copper within the rubblized zone.

Redhawk is currently exploring an area north of Copper Creek which is referred to as Copper Creek North. This area shows the potential for copper mineralization.


SGS FORM No. A263a-2/14/13 !

Redhawk Copper, Inc. - Copper Creek PEA 25,000 TPD Mill Page !24224.0 OTHER RELEVANT DATA AND INFORMATION

The information presented herein is considered to be sufficient for a preliminary economic assessment and SGS is not aware of any other relevant data or information.


SGS FORM No. A263a-2/14/13 !

Redhawk Copper, Inc. - Copper Creek PEA 25,000 TPD Mill Page !24325.0 INTERPRETATION AND CONCLUSIONS

Conceptual process flowsheets and design have been developed for this study. The processing plant is capable of achieving the design criteria of 25,000 tons per day with an overall availability factor of 92.5 percent and the process will yield metal recoveries as stated in the report.

The exploration to date has identified both breccia hosted and porphyry style hosted copper, molybdenum and precious metals mineralization in significant enough quantities to support an underground mining scenario. The mineralization is open in many directions and there are indications though the mapping of the trace elements, those other significant and undiscovered mineralized bodies exist within the Copper Creek property. As well, most of the mapped breccia deposits that are exposed at surface have not been tested with drill holes for the grade of the mineralization. The discovery of additional, high grade mineralization will help the overall project economics and thus Redhawk should continue its exploration program to investigate the opportunities for additional mineralized bodies.


SGS FORM No. A263a-2/14/13 !

Redhawk Copper, Inc. - Copper Creek PEA 25,000 TPD Mill Page !24426.0 RECOMMENDATIONS

The following recommendations are provided by discipline and have not been assigned levels of importance.

26.1 Geotechnical Investigation

A geotechnical investigation will be required in the plant site, the diversion tunnel inlet and outlet areas, and at the proposed TSF site. The focus of the geotechnical investigation will be to develop an understanding of foundation conditions, identify and characterize construction materials, support stability analyses, and meet ADEQ requirements for general site characterization. A program involving test pit excavation and drilling will be required. Costs for a geotechnical site investigation to support detailed engineering and ADEQ permitting will be on the order of $250,000 to $300,000 with testing and subcontractor costs included. Preliminary sites investigation costs for prefeasibility level assessment will be under $100,000.

26.2 Tailings Characterization Studies

A bulk sample of process tailings will be required to support tailings characterization studies. The sample should be passed through a cyclone to produce underflow and overflow samples. Underflow testing should include compressive strength testing of cement amended samples to determine binder addition requirements and costs for hydraulic mine backfill. Tailings overflow and whole tailings should be subjected to geotechnical testing to evaluate consolidation characteristics and post deposition density. Dewatering tests will be required. Subsamples of all tailings products should also be subject to geochemical characterization testing as prescribed by ADEQ. This program will require on the order of 300 kg of tailings solids in order to produce multiple underflow samples for strength testing. Testing program costs, exclusive of the metallurgical testing component (milling, flotation etc), will be on the order of $150,000.

26.2.1 Hydraulic Backfill

In order to further advance the Project, Paste Engineering and Design Group (PED) presents the following key recommendations:

▪ Investigate the unconfined compressive strength (UCS) strength requirement for backfill;

▪ Perform cyclone testwork with a vendor to determine the ideal cyclone size for the process, and obtain cyclone underflow samples in order to perform UCS testwork;

▪ Complete a full laboratory testing program to assess the dewatering potential, rheological and strength properties of the Copper Creek mill tailings for use as cemented underground hydraulic backfill;


SGS FORM No. A263a-2/14/13 !

Redhawk Copper, Inc. - Copper Creek PEA 25,000 TPD Mill Page !245▪ Investigate commercial sources of binder (fly ash, lime kiln dust or other

pozzolans) in the area and evaluate the total cost implications of using partial binder alternatives in place of normal Portland cement; and

Commence pre-feasibility engineering following completion of the above items to determine a greater accuracy cost estimates and process design.

26.3 Alternative Dam Construction Methods

The PEA level cost estimate assumes TSF embankment construction by the downstream raise method. The approach requires the largest amount of fill material relative to other construction methods. Downstream construction will be feasible regardless of the properties of the tailings placed inside the TSF. Rejects from the backfill plant will consist of tailings slimes that could have poor permeability and consolidation characteristics. Poorly consolidated, soft and saturated tailings will not adversely impact TSF stability if the downstream construction method is used, however, utilization of upstream and centerline raise construction techniques, which could reduce embankment fill requirements, is contingent upon the development of well drained and consolidated foundation conditions to support raises. A portion of the testing described in Section 26.2 will include consolidation, settling, permeability, and shear strength testing of backfill plant rejects and whole tailings. The results of these tests should be used in 1-dimensional consolidation modeling to evaluate future conditions on the TSF beach area, and the potential for the use of alternative dam construction practices to reduce fill quantities and construction costs. Consolidation modeling costs will be on the order of $30,000.

26.4 Recommendations for Water Supply Development

For ongoing investigation of the project water supply, drilling and testing of at least one test well is recommended. The well should be located at least three miles east of the active channel of the San Pedro River near the main access road.

A 10-inch well would be physically capable of producing 600 gpm at the estimated depth to groundwater. Upon completion of the well, a limited 10 hour aquifer test would be performed and the results evaluated. Estimated costs for installation and short term testing of the 10- inch test well are contained in Table 20.1.

If data from the 10 hour test are satisfactory then an extended aquifer test should be performed. The purpose of the long term, 15 day test would be to evaluate drawdown associated with long term pumping, which is currently unknown. Costs for long term testing, oversight and evaluation of test data are estimated to be US$ 63,000.

Given a satisfactory long term test, installation of additional 10-inch production wells could be warranted at locations no closer than one quarter mile from the initial 10-inch test well. The combination of the five wells would meet the requirements of the project. The approximate costs for a 10 -inch production well are included in Table 26.1.


SGS FORM No. A263a-2/14/13 !

Redhawk Copper, Inc. - Copper Creek PEA 25,000 TPD Mill Page !246


SGS FORM No. A263a-2/14/13 !

Redhawk Copper, Inc. - Copper Creek PEA 25,000 TPD Mill Page !247Table 26.1

Redhawk Copper Test Wells – Mammoth - Arizona

Item Description Unit

Est. Numbe

r of Units

Unit Price US$

Total Price US$

1 Mobilization and Demobilization Lump Sum 1$25,000.0

0$25,000.0

0

2 Noise Control (optional)Linear Feet 200 $50.00

3Surface Casing Material and Installation (16-inch diameter)

Linear Feet 40 $625.00

$25,000.00

4 Exploration Borehole Drilling (6-inches diameter)Linear Feet 1500 $42.00

$63,000.00

5 Geophysical Logging Lump Sum 1$12,000.0

0$12,000.0

0

6 Packer Testing and Zonal Sampling

A. Furnish, install and remove packer test and ancillary equipment (approximately 3 zones per well) Lump Sum 3 $3,900.00 $11,700.00

B. Airlift water sample for analysis Hourly 24 $650.00$15,600.0

0

C. Conduct slug tests at each interval Hourly 24 $650.00$15,600.0

0

7 Reaming Borehole (14-inches diameter)Linear Feet 1200 $52.00

$62,400.00

8 Casing and Well Screen Material

A. 10-inch diameter LCS Well CasingLinear Feet 500 $54.00

$27,000.00

B. 10-inch diameter LCS Well Screen Wire Wrap Slot 0.080"

Linear Feet 700 $99.00

$69,300.00

9 Casing and Well Screen InstallationLinear Feet 1200 $37.00

$44,400.00

10 Annular Material and Installation

A. Silica Sand Filter Pack (8x12 Mesh)Cubic Yards 18.0 $1,795.00

$32,310.00

B. Formation Stabilizer (3/8-inch pea gravel)Cubic Yards 12.00 $125.00 $1,500.00

C. Bentonite Seal Cubic Feet 7.00 $28.00 $196.00

D. Cement Grout Seal Cubic Feet 1.00 $13.00 $13.00

11 Well Development

A. Swab and Air lift development Hourly 20 $650.00$13,000.0

0

B. Clean Out Casing Lump Sum 1 $3,500.00 $3,500.00

C. Pump and Surge development Hourly 20 $300.00 $6,000.00


SGS FORM No. A263a-2/14/13 !


Development of a water supply sufficient for project uses will be required. This process may be complicated by proximity of identified production well sites to the San Pedro River, which is a protected watershed. Based on local geologic and hydrogeological conditions, a lower, basin-fill confined aquifer is expected to occur at the proposed production well site which represents the targeted water source. Groundwater withdrawal from the confined aquifer is expected to enable Redhawk to withdraw water without surface water impacts to the San Pedro River or watershed. Test well installation and pump testing will be required to demonstrate that surface water resources in the watershed will not be impacted by groundwater withdrawals from the lower basin-fill confined aquifer system.

26.5 Additional Metallurgical Testwork

SGS reviewed the metallurgical testing on the Redhawk project and recommended additional testing for the pre-feasibility study to demonstrate recovery of byproducts and optimize the design of the processing plant. A review of the metallurgical testing should be conducted when all of the results are available. Any additional testing necessary can be identified at that time. Additional reviews of the testing requirements should be conducted as the project advances. Prefeasibility level testing would include sample preparation, comminution, flotation, and support investigations at an approximate total cost of $380,000.

26.6 Resource Expansion

IMC recommends that in addition to the work to advance the project to a Pre-Feasibility level, Redhawk continues to expand the resource base at Copper Creek. The distributions of many of the trace elements have been studied by Redhawk as part of its work to define the mineralizing events at Copper Creek. The Redhawk geologists have developed a model of the trace element distribution in relationship to the Keel- American Eagle copper-moly porphyry. Applying this model to other areas on the Copper Creek claim block, the Redhawk geologists see other potential areas where large porphyry

D. Plumbness and Alignment Test Lump Sum 1 $1,500.00 $1,500.00

E. Final Well Video Lump Sum 1 $1,000.00 $1,000.00

Well Testing

A. Funish, install and remove test pumping, monitoring and ancillary equipment Lump Sum 2

$10,000.00

$20,000.00

B. Pumping Tests (10-hr step test, 24-hour constant) Hourly 68 $250.00$17,000.0

0

12 Surface Completion Lump Sum 2 $1,000.00 $2,000.00

A. With crew Hourly --- $650.00

B. Without crew Hourly --- $750.00

Total$469,019.

00


SGS FORM No. A263a-2/14/13 !

Redhawk Copper, Inc. - Copper Creek PEA 25,000 TPD Mill Page !249bodies could exist. This work needs to continue and be drill tested. Redhawk should further define the program required and develop a budget for it.

26.7 Tailings Storage Facility

26.7.1 Waste Rock TSF Embankment Fill

There is potential to mine the upper portion of the Old Reliable by open pit methods. A trade off study evaluating and optimizing this potential is recommended. The use of mine waste rock for TSF embankment fill could potentially reduce fill costs and decrease disturbance area associated with fill borrowing and waste rock disposal.

26.7.2 TSF Construction Methods

As noted above, the evaluation of future tailings properties could demonstrate the feasibility of alternative construction methods for TSF embankment raises. Upstream or centerline dam raises could reduce embankment fill requirements in later TSF construction phases. Alternative construction methods will have no impact on initial capital costs, however; sustaining capital costs could be significant reduced if fill requirements are reduced.


SGS FORM No. A263a-2/14/13 !

Redhawk Copper, Inc. - Copper Creek PEA 25,000 TPD Mill Page !25027.0 REFERENCES

Arizona Game & Fish Department (AGFD). 1996. Wildlife of Special Concern in Arizona Draft. Non-Game and Endangered Wildlife Program, Arizona Game & Fish Department, Phoenix, Arizona. October 14, 1996.

Bureau of Land Management. 2010. Arizona Bureau of Land Management Sensitive Species List. Instructional Memorandum No. AZ-2011-005, December 22, 2010. BLM Internet Web Site: http://www.blm.gov/pgdata/etc/medialib/blm/az/pdfs/efoia/2011IM_IB.Par.64105.File.dat/IMAZ-2011-005.pdf.

Hart, S., P. Woodman, S. Bailey, S. Boland, P. Frank, G. Goodlett, D. Silverman, D. Taylor, M. Walker and P. Wood. 1992. Desert tortoise population studies at seven sites and a mortality survey at one site in the Sonoran Desert, Arizona. Arizona Game & Fish Department, Phoenix.

Independent Mining Consultants, Inc. report titled "Copper Creek December 2012 Mineral Resources, Pinal County, Arizona, USA Technical Report, dated January 31, 2013.

Murphy R.W., Berry K.H, Edwards T, Leviton A.E., Lathrop A., Riedle, J.D. 2011. The

dazed and confused identity of Agassiz’s land tortoise, Gopherus agassizii (Testudines, Testudinidae) with the description of a new species, and its consequences for conservation. ZooKeys 113: 39–71. doi: 10.3897/zookeys.113.1353

Proffett, J. M., 2009 High grades in porphyry copper deposits and their relationships to emplacement of magmatic sources: Geology, v. 37 p. 675-678

RS Means, Heavy Construction Cost Data, 26th Edition, RS Means, 2012

U.S. Fish & Wildlife Service (USFWS). 1990. Endangered and threatened wildlife and plants; determination of threatened status for the Mojave population of the desert tortoise; final rule. Federal Register 55:12178–12191.

2008. http://www.USFWS.gov/nevada/desert_tortoise/documents/reports/2011/ Chapter%201_Distance%20Sampling%20and%20Desert%20Tortoises_11Feb2010.pdf

2009. Endangered and threatened wildlife and plants; 90-day finding on a petition to list the Sonoran population of desert tortoise (Gopherus agasizzii) [sic] as a Distinct Population Segment (DPS) with Critical Habitat. Federal Register 74:44335–44344.


SGS FORM No. A263a-2/14/13 !

http://www.blm.gov/pgdata/etc/medialib/blm/az/pdfs/efoia/2011IM_IB.Par.64105.File.dat/IMAZ-2011-005.pdf

http://www.USFWS.gov/nevada/desert_tortoise/documents/reports/2011/

Redhawk Copper, Inc. - Copper Creek PEA 25,000 TPD Mill Page !2512010. Endangered and threatened wildlife and plants: 12-month finding on petition to list

the Sonoran desert tortoise as threatened or endangered; proposed rule. Federal

2011. Endangered and Threatened Wildlife and Plants; Review of Native Species That Are Candidates for Listing as Endangered or Threatened; Annual Notice of Findings on Resubmitted Petitions; Annual Description of Progress on Listing Actions; Proposed Rule. Federal Register 76 (207): 66370-66439.

2012a. General Species Information Sonoran desert tortoise (Gopherus morafkai). Accessed November 2012 at: http://www.fws.gov/southwest/es/arizona/Sonoran_Tort.htm

2012b. Species Profile Sonoran desert tortoise (Gopherus morafkai). Accessed November 2012 a t : h t tp : / /ecos. fws.gov/spec iesProf i le /pro f i le /speciesProfile.action?spcode=C07G

U.S. Forest Service. 2007. Southwestern Region Sensitive Animals. USFS Internet Web Site: http://www.fs.usda.gov/detail/r3/plants-animals/?cid=FSBDEV3_022105.

WildEarth Guardians and Western Watersheds Project. 2008. Petition to list the Sonoran desert tortoise (Gopherus agassizii) as threatened or endangered under the Endangered Species Act. Report to the U.S. Department of the Interior, Fish & Wildlife Service.

Woodman, P., P. Frank, and D. Silva. 1999. Desert tortoise population surveys at three sites in the Sonoran Desert of Arizona, 1998. Report to Arizona Game & Fish Department, Phoenix.


SGS FORM No. A263a-2/14/13 !

http://ecos.fws.gov/speciesProfile/profile/speciesProfile.action?spcode=C07G

http://www.fs.usda.gov/detail/r3/plants-animals/?cid=FSBDEV3_022105

Redhawk Copper, Inc. - Copper Creek PEA 25,000 TPD Mill Page !25228.0 APPENDIXES

Appendix Documents and Drawings can be viewed and are available in the Redhawk Copper, Inc.

offices San Manuel, AZ after signing the confidentiality agreement.


SGS FORM No. A263a-2/14/13 !


APPENDIX 1 Redhawk Copper, (USA), Inc. Property List


SGS FORM No. A263a-2/14/13 !


APPENDIX 2 Call and Nicholas, Inc. Report


SGS FORM No. A263a-2/14/13 !


APPENDIX 3 Copper Creek Flowsheet 25ktpd


SGS FORM No. A263a-2/14/13 !


APPENDIX 4 Process Plant Design Criteria


SGS FORM No. A263a-2/14/13 !


APPENDIX 5 Process Plant Flowsheets

APPENDIX 6 Process Plant Equipment List


SGS FORM No. A263a-2/14/13 !


APPENDIX 7A Siting Study for Production Well Locations


SGS FORM No. A263a-2/14/13 !


APPENDIX 7B State Land Well Aquifier Testing


SGS FORM No. A263a-2/14/13 !


APPENDIX 7C Well Spacing Study


SGS FORM No. A263a-2/14/13 !


APPENDIX 8 Process Plant Capital Cost Estimate


SGS FORM No. A263a-2/14/13 !