mina de cobre panamÁ project panamÁ ni 43-101 …mina de cobre panamÁ project, panamÁ ni 43-101...

MINA DE COBRE PANAMÁ PROJECT, PANAMÁ NI 43-101 TECHNICAL REPORT

Prepared for: Prepared by: Inmet Mining Corporation WLR Consulting, Inc. 330 Bay Street, Suite 1000 9386 West Iowa Avenue Toronto, Ontario Lakewood, CO 80232 USA Canada M5H 2S8 Phone: 303-980-8528 Phone: 416-361-6400 Authors and Qualified Persons: William L. Rose, P.E. (Principal Author) Colin Burge, P.Geo. Bruce Davis, F.AusIMM Alexandra Kozak, P.Eng. Robert Sim, P.Geo. Gary S. Wells, P.Geo.

Effective Date: March 31, 2010 Execution Date: May 3, 2010

INMET MINING CORPORATION

MINA DE COBRE PANAMÁ NI 43-101 TECHNICAL REPORT

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T A B L E O F C O N T E N T S

1.0 SUMMARY .................................................................................................................. 1

2.0 INTRODUCTION ....................................................................................................... 10

3.0 RELIANCE ON OTHER EXPERTS ........................................................................... 12 3.1 Terms of Reference ....................................................................................... 12

4.0 PROPERTY DESCRIPTION AND LOCATION ......................................................... 16 4.1 Location ......................................................................................................... 16 4.2 Property Description and Legal Status .......................................................... 16 4.3 Permits .......................................................................................................... 20 4.4 Royalties and Taxes ...................................................................................... 21

5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY ...................................................................................................... 22 5.1 Accessibility ................................................................................................... 22 5.2 Climate .......................................................................................................... 22 5.3 Local Resources and Infrastructure ............................................................... 22 5.4 Physiography ................................................................................................. 23

6.0 HISTORY .................................................................................................................. 24

7.0 GEOLOGICAL SETTING .......................................................................................... 27 7.1 Regional Geological Setting .......................................................................... 27 7.2 Lithology ........................................................................................................ 30 7.3 Alteration ....................................................................................................... 32 7.4 Botija .............................................................................................................. 34 7.5 Colina ............................................................................................................ 37 7.6 Valle Grande .................................................................................................. 39

8.0 DEPOSIT TYPES ...................................................................................................... 42

9.0 MINERALIZATION .................................................................................................... 43 9.1 Introduction .................................................................................................... 43 9.2 Botija .............................................................................................................. 45 9.3 Colina ............................................................................................................ 45 9.4 Valle Grande .................................................................................................. 45 9.5 Brazo ............................................................................................................. 46 9.6 Botija Abajo ................................................................................................... 46 9.7 Medio ............................................................................................................. 46 9.8 Cuatro Crestas .............................................................................................. 47 9.9 Lata ................................................................................................................ 47 9.10 Nada .............................................................................................................. 47

10.0 EXPLORATION ......................................................................................................... 48 10.1 Historical Surveys .......................................................................................... 48

10.1.1 Adrian (1992-1995) ............................................................................ 48 10.1.2 Petaquilla Copper (2006–2008) ......................................................... 49

10.2 MPSA (2007–2009) ....................................................................................... 50

11.0 DRILLING .................................................................................................................. 51 11.1 Historical Drilling ............................................................................................ 52



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11.1.1 United Nations Development Program (1967-1969) .......................... 52 11.1.2 Panamá Mineral Resources Development (PMRD) (Japanese

Consortium) ....................................................................................... 53 11.1.3 Adrian Resources (1992 – 1995) (Inmet-Adrian-Georecursos) ......... 53 11.1.4 Teck (1996–1997) .............................................................................. 53 11.1.5 Petaquilla Copper (2006–2008) ......................................................... 54

11.2 MPSA (2007–2009) ....................................................................................... 54 11.3 Surveying ....................................................................................................... 55

11.3.1 Downhole Surveying .......................................................................... 55 11.3.2 Collar Surveys ................................................................................... 55

12.0 SAMPLING METHOD AND APPROACH ................................................................. 57 12.1 Historical Drilling ............................................................................................ 57

12.1.1 UNDP ................................................................................................. 57 12.1.2 PMRD ................................................................................................ 57 12.1.3 Adrian, Teck ....................................................................................... 57 12.1.4 Petaquilla Copper .............................................................................. 57

12.2 MPSA ............................................................................................................ 57

13.0 SAMPLE PREPARATION, ANALYSES AND SECURITY ........................................ 59 13.1 Historical Drilling ............................................................................................ 59

13.1.1 UNDP ................................................................................................. 59 13.1.2 PMRD ................................................................................................ 59 13.1.3 Adrian, Teck ....................................................................................... 59 13.1.4 Petaquilla Copper .............................................................................. 60

13.2 MPSA ............................................................................................................ 61

14.0 DATA VERIFICATION ............................................................................................... 63 14.1 Historical Drilling ............................................................................................ 63

14.1.1 Adrian, Teck ....................................................................................... 63 14.1.2 Petaquilla Copper .............................................................................. 63

14.2 MPSA ............................................................................................................ 66 14.3 Database Validation ...................................................................................... 77

15.0 ADJACENT PROPERTIES ....................................................................................... 78 15.1 Molejón .......................................................................................................... 78

16.0 MINERAL PROCESSING AND METALLURGICAL TESTING ................................. 79 16.1 Metallurgical Testing ...................................................................................... 79

16.1.1 Introduction ........................................................................................ 79 16.1.2 Samples ............................................................................................. 80 16.1.3 Grindability Testwork ......................................................................... 80 16.1.4 Flotation Variability Testwork ............................................................. 80 16.1.5 2009 Pilot Plant .................................................................................. 80

16.2 Mineral Processing ........................................................................................ 81

17.0 MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES ............................ 84 17.1 Mineral Resource Estimates .......................................................................... 84

17.1.1 Introduction ........................................................................................ 84 17.1.2 Geologic Model, Domains, and Coding ............................................. 85 17.1.3 Available Data .................................................................................... 89 17.1.4 Compositing ....................................................................................... 91 17.1.5 Exploratory Data Analysis .................................................................. 92 17.1.6 Bulk Density Data .............................................................................. 98



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17.1.7 Evaluation of Outlier Grades .............................................................. 99 17.1.8 Variography ..................................................................................... 100 17.1.9 Model Set-Up and Limits ................................................................. 102 17.1.10 Interpolation Parameters .................................................... 103 17.1.11 Validation ........................................................................... 104 17.1.12 Resource Classification ...................................................... 109 17.1.1 Mineral Resource Estimates ............................................................ 111 17.1.2 Comparison with Pevious Estimate ................................................. 115

17.2 Mine Planning & Mineral Reserves ............................................................. 116 17.2.1 Pit Limit Analyses ............................................................................ 116 17.2.2 Recoveries ....................................................................................... 116 17.2.3 Overall Slope Angles ....................................................................... 119 17.2.4 Price Sensitivity Evaluations ............................................................ 119

17.3 Mining Pit/Phase Designs ............................................................................ 122 17.4 Mineral Reserve Estimates .......................................................................... 125

17.4.1 Ore Definition Parameters ............................................................... 125 17.4.2 Material Densities ............................................................................ 127 17.4.3 Dilution ............................................................................................. 127 17.4.4 Ore Recovery ................................................................................... 128 17.4.5 Mineral Reserve Estimate Summary ............................................... 128

18.0 OTHER RELEVANT DATA AND INFORMATION .................................................. 132

19.0 INTERPRETATION AND CONCLUSIONS ............................................................. 133

20.0 RECOMMENDATIONS ........................................................................................... 135

21.0 REFERENCES ........................................................................................................ 137

22.0 DATE AND SIGNATURES ...................................................................................... 139

23.0 ADDITIONAL REQUIREMENTS FOR TECHNICAL REPORTS ON DEVELOPMENT PROPERTIES AND PRODUCTION PROPERTIES ................... 151 23.1 Mining Operations ....................................................................................... 151 23.2 Ore Processing ............................................................................................ 153

23.2.1 Tailings Disposal .............................................................................. 157 23.3 Metallurgical Recoveries and Concentrator Production .............................. 158 23.4 Infrastructure ............................................................................................... 159

23.4.1 Mine/Plant Site ................................................................................. 163 23.4.2 Eastern Infrastructure Area .............................................................. 163 23.4.3 Port Site Infrastructure ..................................................................... 163 23.4.4 Security Buildings ............................................................................ 165 23.4.5 On-Site Roads ................................................................................. 165 23.4.6 Access Roads to Site ....................................................................... 165

23.5 Markets ........................................................................................................ 166 23.5.1 Scope ............................................................................................... 166 23.5.2 Concentrate Production ................................................................... 166 23.5.3 Quality .............................................................................................. 167

23.6 Contracts and Agreements .......................................................................... 169 23.7 Environmental Considerations ..................................................................... 169 23.8 Taxes and Royalties .................................................................................... 172 23.9 Capital Costs ............................................................................................... 173 23.10 Operating Costs ........................................................................................... 176 23.11 Economic Analysis ...................................................................................... 177



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23.12 Payback ....................................................................................................... 181 23.13 Mine Life ...................................................................................................... 181

T A B L E S

Table 1-1: Measured + Indicated Mineral Resources - Mina de Cobre Panamá .............................. 3 Table 1-2: Inferred Mineral Resources - Mina de Cobre Panamá ..................................................... 3 Table 1-3: Mineral Reserve Estimates by Classification and Ore Type ............................................ 4 Table 1-4: Concentrator Production Summary .................................................................................. 6 Table 1-5: Capital Expenditures (millions) ......................................................................................... 8 Table 1-6: Summary of Operating Cost Estimate ($/t milled) ............................................................ 8 Table 1-7: Long-Term Metal Price Assumptions ............................................................................... 9 Table 2-1: Dates of Site Visits and Areas of Responsibility ............................................................ 11 Table 4-1: MPSA Mineral Concessions under Law No. 9, 1997 ..................................................... 20 Table 6-1: Exploration History of the Cobre Panamá Concession .................................................. 25 Table 11-1: Summary of Drilling by Operator and Area .................................................................... 51 Table 11-2: Summary of Drilling by Area ........................................................................................... 51 Table 15-1: 2007 Resource Estimate – Molejón Gold Deposit (Petaquilla Minerals) ....................... 78 Table 16-1: Concentrator Production Summary ................................................................................ 80 Table 17-1: Distribution of Drilling Data by Area ............................................................................... 90 Table 17-2: Statistical Summary of Sample Assay Data by Area ..................................................... 91 Table 17-3: Summary of Interpolation Domains ................................................................................ 98 Table 17-4: Summary of Bulk Density Values by Rock Type ............................................................ 98 Table 17-5: Summary of Outlier Grade Controls ............................................................................... 99 Table 17-6: Variogram Parameters ................................................................................................. 100 Table 17-7: Block Model Limits ........................................................................................................ 102 Table 17-8: Interpolation Parameters .............................................................................................. 104 Table 17-9: Summary of Measured Mineral Resources .................................................................. 111 Table 17-10: Summary of Indicated Mineral Resources ................................................................... 111 Table 17-11: Summary of Measured And Indicated Mineral Resources ........................................... 113 Table 17-12: Summary of Inferred Mineral Resource ....................................................................... 114 Table 17-13: Historical Resource Estimate – 14 January 1998 ........................................................ 115 Table 17-14: Overall Slopes for Floating Cone Analyses .................................................................. 119 Table 17-15: Floating Cone Price Sensitivity Analyses (combined results from Botija, Colina and

Valle Grande) ............................................................................................................... 121 Table 17-16: Basic Pit Design Parameters ........................................................................................ 122 Table 17-17: Pit Slope Design Parameters ....................................................................................... 123 Table 17-18: Ore Definition Parameters for Mineral Reserve Estimates .......................................... 126 Table 17-19: NSR Cutoff Grades (US$/t) used in Mineral Reserve Estimates ................................. 127 Table 17-20: Material Densities used in Mineral Reserve Estimates ................................................ 127 Table 17-21: Proven Mineral Reserve Estimates by Ore Type ......................................................... 128 Table 17-22: Probable Mineral Reserve Estimates by Ore Type ...................................................... 129 Table 17-23: Combined Proven and Probable Mineral Reserve Estimates by Ore Type ................. 130 Table 23-1: Mine Production Schedule ........................................................................................... 151 Table 23-2: Mill Feed Schedule (all ore types) ................................................................................ 153 Table 23-3: Concentrator Production Summary .............................................................................. 159 Table 23-4: Mining Production Parameters ..................................................................................... 167 Table 23-5: Analysis of Copper (Cu) Concentrate ........................................................................... 168 Table 23-6: Analysis of Molybdenum (Mo) Concentrate ................................................................. 169 Table 23-7: Summary of Capital Costs ............................................................................................ 175



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Table 23-8: Summary of Operating Cost Estimate ($/t milled) ........................................................ 176 Table 23-9: Long-Term Metal Price Assumptions ........................................................................... 177 Table 23-10: Summary of Key Financials (base case) ...................................................................... 178 Table 23-11: Impact of Metal Price Change ...................................................................................... 179 Table 23-12: Levered Case Financial Results ................................................................................... 180 Table 23-13: Summary of Base Case Performance Statistics .......................................................... 180 Table 23-14: Summary of Levered Case Performance Statistics ...................................................... 180

F I G U R E S

Figure 1-1: Regional Geology and Deposit Locations ........................................................................ 2 Figure 4-1: Location Map .................................................................................................................. 16 Figure 4-2: General Arrangement of Project Facilities ...................................................................... 18 Figure 4-3: Property Location Map – Mina de Cobre Panamá Concessions ................................... 19 Figure 7-1: Generalized Geology of Panamá ................................................................................... 27 Figure 7-2: District Geology and Structural Lineaments ................................................................... 29 Figure 7-3: Property Geology ............................................................................................................ 30 Figure 7-4: Botija Deposit - Plan Map - Geology, Drill Hole Locations ............................................. 35 Figure 7-5: Section 53 8200 East - Botija Deposit ............................................................................ 36 Figure 7-6: Colina Deposit - Plan Map - Geology, Drill Hole Locations ............................................ 37 Figure 7-7: Section 53 3800 East - Colina Deposit ........................................................................... 38 Figure 7-8: Valle Grande Deposit - Plan Map - Geology, Drill Hole Locations ................................. 40 Figure 7-9: Section 13 West - Valle Grande Deposit ........................................................................ 41 Figure 9-1: Location Map of Mineralized Zones ............................................................................... 44 Figure 10-1: Soil Geochemistry .......................................................................................................... 49 Figure 10-2: Plan Map – IP Chargeability at -150 meter Level ........................................................... 50 Figure 11-1: Drill Hole Location Map .................................................................................................. 52 Figure 14-1: Results for ORE 52Pb – Reference Material Certified for Cu ........................................ 64 Figure 14-2: Results for ORE 53 Pb - Reference Material Certified for Au ........................................ 65 Figure 14-3: Blank Material - Cu% ...................................................................................................... 65 Figure 14-4: Blank Material - Au ppm ................................................................................................. 66 Figure 14-5: Results for CM-1 - Reference material certified for Cu .................................................. 68 Figure 14-6: Results for CM-1 - Reference Material certified for Au .................................................. 69 Figure 14-7: Blank Material - Cu% ...................................................................................................... 70 Figure 14-8: Blank Material - Au g/t .................................................................................................... 71 Figure 14-9: Coarse Reject Duplicate Results - Cu% ......................................................................... 72 Figure 14-10: Coarse Reject Duplicate Results - Au g/t ....................................................................... 73 Figure 14-11: Interlaboratory Checks - Cu% ........................................................................................ 74 Figure 14-12: Interlaboratory Checks - Au g/t ....................................................................................... 76 Figure 16-1: Process Flow Diagram ................................................................................................... 82 Figure 17-1: Drill Hole Plan Map ......................................................................................................... 85 Figure 17-2: Lithology Model Isometric ............................................................................................... 86 Figure 17-3: Botija Lithology Model Isometric ..................................................................................... 87 Figure 17-4: Colina and Valle Grande Lithology Model Isometric ...................................................... 88 Figure 17-5: Botija Boxplot for Copper by Rock Type ........................................................................ 94 Figure 17-6: Colina Boxplot for Copper by Rock Type ....................................................................... 94 Figure 17-7: Valle Grande Boxplot for Copper by Rock Type ............................................................ 95 Figure 17-8: Botija Contact Profile for Copper between Andesite and Granodiorite ......................... 96 Figure 17-9: Botija Contact Profile for Copper between Andesite and Porphyry .............................. 97 Figure 17-10: Block Model Limits ........................................................................................................ 103



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Figure 17-11: Herco Plots for Copper ................................................................................................. 106 Figure 17-12: Grade/Tonnage Comparison of Copper Models .......................................................... 107 Figure 17-13: Swath Plots for Copper................................................................................................. 108 Figure 17-14: Plan Limit of Resource Classes ................................................................................... 110 Figure 17-15: Molybdenum Recoveries by Grade .............................................................................. 118 Figure 17-16: Ultimate Pit and Waste Rock Storage Facility Designs ................................................ 124 Figure 23-1: Process Flow Diagram ................................................................................................. 156 Figure 23-2: Layout of Process Plant Area ....................................................................................... 157 Figure 23-3: Layout of Mine/Plant Site ............................................................................................. 160 Figure 23-4: Layout of Eastern Infrastructure Area .......................................................................... 161 Figure 23-5: General Arrangement of Port Site Facilities ................................................................. 162 Figure 23-6: Sensitivity of After-Tax NPV @ 8% .............................................................................. 178 Figure 23-7: Sensitivity Spider Graph for After-Tax IRR ................................................................... 179



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1.0 SUMMARY

Introduction

This Technical Report was compiled by WLR Consulting, Inc. for Inmet Mining Corporation to provide updated mineral resource and reserve estimates and the mining plan for the Cobre Panamá project. The report was written under the direction of William Rose, P.E., with contributions from Robert Sim, P.Geo., Bruce Davies, FAusIMM., Alexandra Kozak, P.Eng., Colin Burge, P.Geo. and Gary Wells, P.Geo., all Qualified Persons as defined by NI 43-101 and as described in Section 22. This Technical Report summarizes the findings of the Mina de Cobre Panamá Project FEED Study Report completed by AMEC AMERICAS LIMITED (AMEC) on March 31, 2010.

Property Location and Ownership

The Mina de Cobre Panamá Cu-Mo-Au project is located in the Donoso District of Colón province, Republic of Panamá, approximately 120 km west of Panamá City. The center of the project area occurs at latitude 8o50’ North and longitude 80o38’ West. The project is accessible from Panamá City via the Pan-American highway and secondary paved and gravel roads.

Inmet Mining Corporation, through its Panamanian subsidiary Minera Panamá S.A. (MPSA) is 100% owner of the Mina de Cobre Panamá project. The property consists of four concessions totalling 13,600 hectares.

History

Copper-gold-molybdenum porphyry style mineralization was discovered in central Panamá during a regional survey by the United Nations in 1968. Exploration by several companies has since outlined three large deposits and several smaller ones. Companies that have conducted drill programs include: United Nations Development Program (1968-1969), Panamá Mineral Resources Development Company (PMRD), a Japanese consortium (1970-1980), Inmet - Adrian Resources – Teck (formerly Teck Cominco) (1990-1997), Petaquilla Copper (2006-2008) and Minera Panamá S.A. (MPSA) (2007-2009). A total of 1,275 diamond drill holes (230,555 m) have been completed.

Geology

The porphyry deposits occur at the southern margin of a large granodioritic batholith of mid-Oligocene age (36.4 Ma). They occur in a WNW-ESE oriented zone with dimensions of 9 km by 4.5 km (Figure 1-1). The three main deposits are Botija, Colina and Valle Grande. There are also a number of smaller zones; the most significant being Brazo and Botija Abajo.

All of the porphyry style mineralization on the property is hosted in granodiorite, feldspar-quartz-hornblende porphyry and adjacent andesitic volcanics. At Botija, a number of north dipping feldspar-quartz-hornblende dikes cut the granodiorite. Two roof pendants of andesitic volcanics occur in the central and eastern parts of the deposit. At Colina, mineralization is associated with an east-southeasterly trending, shallow north dipping, 2.5 km by 1 km feldspar-quartz-




hornblende porphyry sill and dike complex that intrudes granodiorite and andesitic volcanics. The Valle Grande zone is associated with a southeast trending feldspar-quartz-hornblende porphyry lopolith that is bounded to the north and south by andesitic volcanics and minor granodioritic dikes. At Brazo and Botija Abajo the host rock is dominantly feldspar-quartz or feldspar-quartz-hornblende porphyry.

Hydrothermal alteration along the Cobre mineral trend is primarily silica-chlorite which is interpreted to be a form of propylitic alteration. Potassic alteration, consisting of salmon coloured potassium feldspar and secondary biotite is seen in the central parts of Botija. Argillic and phyllic alteration is patchy in the three main deposits with the latter variety being most prevalent near the tops of the deposits. At Brazo, pervasive sericite, clay and pyrite is associated with well-developed quartz stockworks.

Hypogene sulphides occur as disseminations, micro-veinlets, fracture fillings, and quartz-sulphide stockworks. Chalcopyrite is the dominant copper mineral with lesser bornite. Traces of molybdenite are commonly found in quartz veinlets. There is no significant zone of supergene enrichment at Botija, Colina and Valle Grande. At Brazo, supergene mineralization consisting of chalcocite-coated pyrite and rare native copper occurs to a depth of at least 150 meters.

Figure 1-1: Regional Geology and Deposit Locations




Mineral Resources

The mineral resources for the porphyry deposits on the Cobre Panamá concession were estimated using the entire drill hole and assay database that exists for the project, a pit shell which is defined with a copper price of $2.30 per lb and a Cu cut-off grade of 0.15%. These are presented in Table 1-1 (Measured plus Indicated) and Table 1-2 (Inferred). Mineral resources, that are not mineral reserves, do not have demonstrated economic viability. It cannot be assumed that all or any part of an Inferred mineral resource will be upgraded to an Indicated or Measured mineral resource as a result of continued exploration.

Table 1-1: Measured + Indicated Mineral Resources - Mina de Cobre Panamá

Deposit Category Tonnes (million)

Cu (%)

Mo (%)

Au g/t

Ag g/t

Botija Measured 261 0.56 0.009 0.13 1.50

Indicated 907 0.33 0.007 0.06 1.00

Colina Indicated 1,178 0.35 0.007 0.05 1.50

Valle Grande Indicated 671 0.34 0.006 0.04 1.30

Botija Abajo Indicated 184 0.28 0.004 0.09 0.90

Brazo Indicated 71 0.43 0.004 0.12 0.70

All Areas Measured 261 0.56 0.009 0.13 1.50

Indicated 3,010 0.34 0.006 0.06 1.20

Measured + Indicated 3,271 0.36 0.007 0.06 1.30

Table 1-2: Inferred Mineral Resources - Mina de Cobre Panamá

Deposit Tonnes (million) Cu (%) Mo (%) Au g/t Ag g/t

Botija 407 0.21 0.004 0.03 0.70

Colina 1,090 0.24 0.005 0.03 1.20

Valle Grande 1,141 0.24 0.005 0.03 1.00

Botija Abajo 287 0.22 0.005 0.07 0.90

Brazo 269 0.27 0.004 0.07 0.60

Total All Areas 3,194 0.24 0.005 0.04 1.00




Mineral Reserves

The mine plan was initially developed from the deposit models produced by MPSA and Inmet Mining as of the 1st of November 2009. A series of floating cone analyses were conducted to determine economic pit limits and the mining phase development sequence for three mineral deposits in the project Concession area: Botija, Colina, and Valle Grande. Botija is immediately northeast of the proposed concentrator site, Colina is about 2 km to the west-northwest, and Valle Grande is roughly 1 km to the southwest. The Botija deposit model was subsequently updated in January 2010 to include in-fill drilling that improved the resource classifications. The January 2010 Botija and November 2009 Colina/Valle Grande models were used to estimate mineral reserves.

The floating cone evaluations, mine design, and reserve estimates are based on metal prices of $2.00/lb Cu, $12.00/lb Mo, $750/oz Au, and $12.50/oz Ag. Recoveries for Cu, Mo, and Au vary by grade, ore type, and deposit, while Ag recovery is generally fixed, except for deductions for saprock ore. Over the life of the project, concentrator recoveries will average about 86% for Cu, 59% for Mo, 54% for Au, and 46% for Ag. Weighted average mining costs of $1.33/t were used in the pit limit analyses, along with base ore processing and general & administration (G&A) costs of $3.88/t and $1.49/t, respectively. Grinding rates will vary by ore type and deposit, which affects the unit ore processing and G&A costs and, therefore, the cutoff grades used for reserve estimation. The Mina de Cobre Panamá project mineral reserve estimates are based on proven and probable ore; all inferred mineral resources were treated as waste.

Table 1-3 presents the mineral reserves by classification and ore type, based on cutoff grades that vary by ore type, deposit, and the time period in which the reserves are to be mined. Total proven and probable mineral reserves are estimated at 2,143 Mt grading 0.41% Cu, 0.008% Mo, 0.07 Au g/t and 1.43 Ag g/t. Contained metal is projected at approximately 19.6 billion pounds of copper, 361 million pounds of molybdenum, 4.96 million troy ounces of gold, and 98.7 million troy ounces of silver. Total material within the designed ultimate pits is estimated at 3.444 billion tonnes, resulting in a stripping ratio of 0.61:1 (tonnes of waste per tonne of ore).

Table 1-3: Mineral Reserve Estimates by Classification and Ore Type

Ore Type kt NSR $/t Cu % Mo % Au g/t Ag g/t

Proven Mineral Reserves

Saprock 4,400 11.78 0.43 0.013 0.11 1.62

Andesite 19,500 16.67 0.45 0.010 0.09 1.05

Porphyry 117,100 24.26 0.64 0.010 0.16 1.74

Granodiorite 104,300 21.25 0.57 0.010 0.13 1.57

Total 245,300 22.15 0.59 0.010 0.14 1.61

Probable Mineral Reserves

Saprock 56,100 9.63 0.39 0.007 0.09 1.49

Andesite 414,700 13.52 0.39 0.007 0.06 1.42

Porphyry 927,500 14.50 0.41 0.007 0.07 1.52

Granodiorite 499,100 12.55 0.36 0.007 0.05 1.20

Total 1,897,400 13.63 0.39 0.007 0.06 1.41




Proven + Probable Mineral Reserves

Saprock 60,500 9.79 0.39 0.008 0.09 1.50

Andesite 434,200 13.66 0.39 0.007 0.06 1.40

Porphyry 1,044,500 15.59 0.44 0.008 0.08 1.54

Granodiorite 603,400 14.05 0.39 0.008 0.07 1.26

Total 2,142,600 14.60 0.41 0.008 0.07 1.43

Note: Estimates based on metal prices of $2.00/lb Cu, $12.00/lb Mo, $750/oz Au, $12.50/oz Ag, and variable NSR cutoff grades

The above mineral reserve projections are contained within the estimates of measured and indicated mineral resources (see Table 1-1). Mr. William Rose, P.E., Principal Mining Engineer of WLR Consulting, Inc. estimated the mineral reserves, which have an effective date of March 31, 2010. Mr. Rose meets the requirements of an independent Qualified Person under the standards of NI 43-101.

Metallurgy

The property has been investigated on behalf of several owners since 1968, and preliminary feasibility studies and prefeasibility studies were done in 1977, 1979, and 1994; feasibility studies were produced in 1994 (updated in 1995), 1996, and 1998. In all of these studies, testwork was done commensurate with the requirements of the times; the study produced in 1997 and published in early 1998 (Teck Corporation Petaquilla Feasibility Study, Simons Project No. U11G, Volume 1, January 1998) built mostly upon work done in the earlier studies.

In 1997, an extensive program of metallurgical testing was designed to confirm earlier work on the metallurgical response of the Botija and Colina deposits. Most of the work was done at Lakefield Research Ltd., Lakefield, Ontario. Grinding, flotation, dewatering, and mineralogical work were performed as part of this program. In addition to the Lakefield work, locked-cycle flotation testwork and modal analysis were performed at G&T Metallurgical Services Ltd., Kamloops, B.C. (G&T) to assist in defining grind requirements for both rougher and cleaner flotation. Copper-molybdenum separation using differential flotation was conducted by International Metallurgical and Environmental, Kelowna, B.C. (IME). The metallurgical work done for the present study has built upon the 1997/1998 study with some knowledge of, but no reliance on, work performed before that time.

The testwork before 2007 was based on large composite samples, and the results, particularly for flotation testing, could not be used for interpreting the variability of response for material within the deposits. Consequently, a large sampling program was undertaken in 2008/2009 to bolster the knowledge from previous work and provide the missing insight into the variability of response. A total of 16 special holes for metallurgical grinding and flotation tests were drilled in the Botija, Valle Grande, and Colina orebodies. Sample preparation, flotation testing, and testing of flotation products were done primarily at G&T Metallurgical Services in Kamloops, British Columbia. Grinding work was conducted at SGS Mineral Services in Lakefield, Ontario, and at Philips Enterprises LLC in Golden, Colorado.




The program resulted in:

additional geological data a comprehensive suite of grindability parameters resulting in new throughput

estimates additional flotation response data for estimating concentrates production and operating

costs sample materials for marketing purposes additional design data for solid-liquid separation, regrinding, and pipeline design.

The resulting life-of mine production data are shown in Table 1-4.

Table 1-4: Concentrator Production Summary

Parameter Unit Years 2-20 Years 21-30 Life of Mine

Throughput tonnes 1,367,393,000 732,814,000 2,142,652,000

t/a 71,968,053 73,281,400 71,421,733

t/d 197,173 200,771 195,676

Head Grade % Cu 0.46 0.34 0.41

% Mo 0.008 0.006 0.008

g/t Au 0.08 0.05 0.07

oz/t Ag 1.46 1.39 1.43

Recovery % Cu 88.4 79.5 85.9

% Mo 61.9 53.1 59.0

% Au 57.2 44.6 54.3

% Ag 47.3 42.8 45.8

Copper Concentrate Production t/a 1,033,685 697,813 909,625

% Cu 28 28 28

Molybdenum Concentrate Production

t/a 7,090 4,741 6,188

% Mo 52 52 52

Ore Processing

Ore from the Botija, Colina, and Valle Grande pits will be treated in a concentrator to produce a copper concentrate and a molybdenum concentrate for sale on the world market. Initially, the concentrator will treat nominally 150,000 t/d of ore supplied from the Botija pit; later, ore will be received from the Colina and Valle Grande pits. From Year 10, the concentrator ore throughput will be increased by 50%, to a nominal 225,000 t/d, to maintain production of concentrate despite a falling head grade. Crushing, grinding, bulk rougher flotation, water, and air systems will increase in capacity by 50% to accomplish the increase in ore treatment rate; all other systems will remain at the same size.

The process plant is designed to process ore at a head grade of 0.70% Cu and 0.013% Mo. These levels are higher than the highest sustained head grades of 0.58% Cu and 0.011% Mo expected to be mined in Year 5, but the design provides the flexibility to accommodate a wide




range of head grades over the project life. The plant design also allows for 15% day-to-day fluctuations in throughput. The process includes the following facilities:

crushing and grinding to liberate minerals from the ore froth flotation to separate most of the copper and molybdenum minerals from

minerals of no commercial worth differential flotation to separate the copper and molybdenum minerals from each

other facilities to store tailings and provide reclaim water for the process facilities to remove water from the products and to ship concentrates to market.

Infrastructure

There will be several distinct and separate construction areas:

mine site, including the process plant site and camp facilities port site and camp facilities 300 MW power plant (construction and operation by Suez) tailings management facility (TMF) and associated infrastructure Botija pit initially followed by the Colina and Valle Grande pits 230 kV overhead power line Eastern Access Road and upgrades from Llano Grande Coast Road and pipelines.

Project Capital

The total estimated cost to design and build the Mina de Cobre Panamá 150,000 t/d project described in this report is US$4,320 million, including an Owner-provided mining fleet and self-performed preproduction development. This amount covers the direct field costs of executing the project, plus the Owner’s indirect costs associated with design, construction, and commissioning. The estimate is summarized in Table 1-5,

All costs are expressed in second quarter 2009 (Q2) US dollars. No allowance has been included for escalation, interest, financing fees, taxes, duties, or working capital during construction. The level of accuracy for the estimate is ±15% of estimated final costs.




Table 1-5: Capital Expenditures (millions)

Mining US$510

Process plant 799

Site and services 597

Port site 320

Total direct cost 2,227

Construction indirect 805

Owner’s cost 364 Engineering, procurement and construction management costs 472

Contingency 453

TOTAL CAPITAL COST US$4,320

Operating Costs

The operating cost estimate (opex) has been prepared as an annual cost for the project from plant start-up to mine closure. The life of the project is 30 years at a nominal processing plant throughput rate of 150,000 t/d (54.8 Mt/a) for the first nine years, followed by an increase to a nominal throughput of 225,000 t/d (82.1 Mt/a) from Year 10 onward.

The operating cost estimate is expressed in constant second quarter 2009 (Q2 2009) U.S. dollars with no allowances for escalation or fluctuation in exchange rates. Costs incurred before plant start-up on Q4 2015 are treated as capital expenditures (capex).

The operating costs are grouped into four cost centres:

Open pit mining Processing Site services General and administration (G&A).

The average costs for the project over the mine life are shown in Table 1-6.

Table 1-6: Summary of Operating Cost Estimate ($/t milled)

Cost Centre Labour Material Power Other Total

Open Pit Mining 0.16 1.69 0.06 0.22 2.14

Processing 0.11 1.44 2.12 0.05 3.72

Site Services 0.12 0.18 0.06 0.38 0.73

G&A 0.09 0.00 0.02 0.52 0.64

Total 0.48 3.31 2.26 1.18 7.23




The total operating cost is $15,490 million over the LOM for a milled feed of 2,143 Mt. The overall unit operating cost is $7.23/t of milled ore.

Economics

Engineering studies have demonstrated the technical feasibility of producing significant quantities of copper, molybdenum, silver, and gold from the Mina de Cobre Panamá project. The economic viability of the project has been evaluated by using a combination of pre-tax and after-tax cash flow analyses, based on the engineering studies and cost estimates discussed herein. Under the metal price assumptions shown in Table 1-7 and using a discount rate of 8%, the pre-tax project net present value (NPV) for the base case is $1,661 million, and the internal rate of return (IRR) is 12.6%. The after-tax NPV is $1,536 million with an IRR of 12.4%. The cumulative pre-tax undiscounted cash flow value for the project is $12,873 million and the payback period is 5.9 years.

Another case (levered) assuming $2.16 billion of debt financing, representing 50% of the preproduction capital, was evaluated. This resulted in a project after-tax IRR of 15.1% – a net improvement of 2.7% over the un-levered base case.

Table 1-7: Long-Term Metal Price Assumptions

Metal Unit Price

Cu US$/lb 2.10

Au US$/oz 885.00

Mo US$/lb 13.00

Ag US$/oz 13.50

All monetary amounts are presented in United States dollars (US$).

For the sake of discounting, cash flows are assumed to occur at the end of each period. All cash flows are discounted to the beginning of Q4 2010.




2.0 INTRODUCTION

AMEC Americas Limited (AMEC) was retained in early 2007 to develop the Draft Interim FEED Report (AMEC, 2008), a study produced for Teck to bring the Mina de Cobre Panamá project (then known as the Petaquilla project) to full feasibility level. AMEC was subsequently commissioned by Inmet Mining, through its wholly-owned subsidiary Minera Panamá S.A. (MPSA), to complete the Front-End Engineering and Design (FEED) Study, which incorporates the results of additional exploration drilling and metallurgical testing.

The purpose of the FEED Study is to describe the status of the Mina de Cobre Panamá project, in the Donoso District of Panamá, in sufficient detail for MPSA to pursue international financing options and move on to the next stage of project engineering and development. One of the study objectives is to produce a capital cost estimate with an accuracy of +15%/-15%, with a confidence level of 80%, of estimated final cost based on 10% completion of engineering.

The results of the FEED program were announced by Inmet Mining in a press release dated March 31, 2010. This technical report is a summary of the FEED report and was prepared in accordance with the requirements of NI 43-101, Companion Policy 43-101CP, and Form 43-101F1 of the Canadian Securities Administrators (CSA) under the direction of William Rose, P.E. with WLR Consulting, Inc., an independent “Qualified Person” as defined by the NI 43-101. Mr. Rose visited the site between April 15th to 16th, 2009 to review the geologic and physiographic settings for the project and to review the data collection program. In addition, the following Qualified Persons responsible for preparing parts of this technical report visited the site on the dates noted. Please refer to section 22 for the Statement of Qualified Person for each Qualified Person named in this report.

Colin Burge, P.Geo., is the site technical manager for the FEED drill program and has been on site since October 2007.

Gary Wells, P.Geo., compiled the geological and drillhole databases and prepared the geological models used in the resource calculations. He visited the site on several occasions since 2008, the most recent being November 14th to November 17th, 2009.

Bruce Davis, F.AusIMM, reviewed the QA/QC for the FEED drill program. He visited the site between April 15th to April 16th, 2009.

Robert Sim, P.Geo., estimated the mineral resources and visited the site between April 15th to April 16th, 2009

Alexandra Kozak, P.Eng., supervised the engineering for Section 16.0, Mineral Processing and Metallurgical Testing, as well as the mineral processing aspects of Section 23.0. She did not visit the site.

Table 2-1 summarizes the responsibilities of the Qualified Persons who compiled this technical report.




Table 2-1: Dates of Site Visits and Areas of Responsibility

QP Name Site Visit Date Area of Responsibility

William Rose 15-16 April 2009 Sections 1, 2, 3, 17.2, 17.3, 17.4, 18, 19, 20, 22, 23.1, 23.4, 23.5.1, 23.6, 23.7, 23.8, 23.9, 23.10, 23.11, 23.12 and 23.13. Overall responsibility for technical report.

Colin Burge Numerous visits Sections 7, 8, 9, 10, 11, 12, 13 and those portions of the summary, conclusions and recommendations that pertain to those sections.

Gary Wells Numerous visits Sections 4, 5, 6, 7, 8, 9, 10, 11, 12, 13, 15, 21 and those portions of the summary, conclusions and recommendations that pertain to those sections.

Bruce Davis 15-16 April 2009 Sections 14, 17.1 and those portions of the summary, conclusions and recommendations that pertain to those sections.

Rob Sim 15-16 April 2009 Section 17.1 and those portions of the summary, conclusions and recommendations that pertain to that section.

Alexandra Kozak No site visit Sections 16, 23.2, 23.3, 23.5.2, 23.5.3 and those portions of the summary, conclusions and recommendations that pertain to those sections.

The information and data obtained for this report came from the FEED Study report by AMEC. Other technical data, documents and reports on the property were also used for this report and are listed in the References (Section 21).

The effective date for the mineral resource and mineral reserve estimates is March 31, 2010.




3.0 RELIANCE ON OTHER EXPERTS

The report was prepared by WLR Consulting, Inc. (WLRC) under the direction of William Rose, P.E., Principal Mining Engineer, a “Qualified Person” for the purposes of NI 43-101 and fulfills the requirements of an “Independent Qualified Person”. The information, conclusions, opinions and estimates contained herein are based on:

The Mina de Cobre Panamá Project FEED Study Report issued by AMEC on March 31, 2010 (the “FEED Study”).

Data, reports and other information supplied by Inmet Mining Corporation and other third party sources.

For the purpose of this report, WLRC has relied on the ownership data (mineral, surface and access rights) and information provided by Inmet Mining Corporation and believes that such data and information were essentially complete and correct to the best of our knowledge and that no information was intentionally withheld. WLRC has not researched property title or mineral rights for the Mina de Cobre Panamá project and expresses no legal opinion as to the ownership status of the property. WLRC has relied upon Dr. Les Hulett of Les Hulett Consulting and Dr. Craig Ford, VP of People and Environment for Inmet Mining, for data and information regarding environmental and socio-political issues facing the project. Dr. Hulett is independent of Inmet Mining, while Dr. Ford is not independent. Mr. Diomedes Gonzalez provided data and information regarding the permits required for project construction and operation. Mr. Gonzalez is Permits & Government Relations Manager for Minera Panamá S.A. and is not independent of Inmet Mining. AMEC , based in Vancouver, Canada, performed engineering studies and designs for the ore processing, port and other support facilities; prepared estimates of the capital and operating costs (excluding mining and mine equipment capital); and performed financial evaluations of the project. WLRC has relied upon these data and information for the preparation of this technical report. AMEC is independent of Inmet Mining.

3.1 Terms of Reference

The FEED Study addresses the mineral resource, mine plan, processing and support facilities, management of tailings, water, and waste rock, site access, transportation of materials and equipment, port requirements, power supply, environmental aspects, and project execution. MPSA commissioned AMEC to prepare this report with input from MPSA, Inmet, and their various consultants, as listed below.

Inmet/MPSA WLR Consulting, Inc.




Sim Geological Inc. BD Resource Consulting, Inc. AMEC AMERICAS LIMITED (AMEC) SGS Mineral Services Limited (Lakefield, Ontario) G&T Metallurgical Services Ltd. Sandwell Inc. Swiss Energy LLC Golder Associates Ltd. Pipeline Systems Inc. (PSI) DJB Consultants Inc. Estudios Electricos

All measurement units used in this report are metric and currency is expressed in US dollars unless otherwise stated.

Units of Measure

Above mean sea level .......................................................................................................................... amsl Ampere ......................................................................................................................................................A Annum (year) ............................................................................................................................................. a Barrel ...................................................................................................................................................... bbl Billion (year) .............................................................................................................................................. G Billion years ago ...................................................................................................................................... Ga Centimetre .............................................................................................................................................. cm Cubic centimetre .................................................................................................................................... cm3 Cubic metre .............................................................................................................................................. m3

Day ............................................................................................................................................................ d Days per week ...................................................................................................................................... d/wk Days per year (annum) ........................................................................................................................... d/a Decibel adjusted .................................................................................................................................... dBa Decibel ..................................................................................................................................................... dB Degree ....................................................................................................................................................... ° Degrees Celsius ....................................................................................................................................... °C Diameter .................................................................................................................................................... ø Dry metric ton ......................................................................................................................................... dmt Gigajoule ................................................................................................................................................. GJ Gram .......................................................................................................................................................... g Grams per litre ........................................................................................................................................ g/L Grams per tonne ...................................................................................................................................... g/t Greater than ............................................................................................................................................... > Hectare (10,000 m2) ................................................................................................................................. ha Hertz ........................................................................................................................................................ Hz Hour ........................................................................................................................................................... h Hours per day ......................................................................................................................................... h/d Hours per week ..................................................................................................................................... h/wk Hours per year ........................................................................................................................................ h/a Horsepower .............................................................................................................................................. hp Joule .......................................................................................................................................................... J




Joules per kilowatt-hour ..................................................................................................................... J/kWh Kelvin .........................................................................................................................................................K Kilo (thousand) ........................................................................................................................................... k Kilocalorie .............................................................................................................................................. kcal Kilogram ................................................................................................................................................... kg Kilograms per cubic metre .................................................................................................................. kg/m3 Kilograms per hour ................................................................................................................................ kg/h Kilograms per square metre ................................................................................................................ kg/m2 Kilojoule ................................................................................................................................................... kJ Kilometre ................................................................................................................................................. km Kilometres per hour .............................................................................................................................. km/h Kilonewton ............................................................................................................................................... kN Kilopascal ............................................................................................................................................... kPa Kilovolt ..................................................................................................................................................... kV Kilovolt-ampere ..................................................................................................................................... kVA Kilowatt ................................................................................................................................................... kW Kilowatt hour ......................................................................................................................................... kWh Kilowatt hours per tonne ..................................................................................................................... kWh/t Kilowatt hours per year ...................................................................................................................... kWh/a Kilowatts adjusted for motor efficiency .................................................................................................. kWe Less than ................................................................................................................................................... < Litre ............................................................................................................................................................ L Litres per minute ................................................................................................................................. L/min Megabytes per second ......................................................................................................................... Mb/s Megapascal ........................................................................................................................................... MPa Megavolt-ampere ................................................................................................................................. MVA Megawatt ............................................................................................................................................... MW Metre ......................................................................................................................................................... m Metres above sea level ....................................................................................................................... masl Metres per minute .............................................................................................................................. m/min Metres per second ................................................................................................................................. m/s Metric ton (tonne) ........................................................................................................................................ t Micrometre (micron) ................................................................................................................................ µm Microsiemens (electrical) ......................................................................................................................... µs Milliamperes ............................................................................................................................................ mA Milligram .................................................................................................................................................. mg Milligrams per litre ................................................................................................................................ mg/L Millilitre .................................................................................................................................................... mL Millimetre ................................................................................................................................................ mm Million ........................................................................................................................................................ M Million tonnes ........................................................................................................................................... Mt Minute (time) .......................................................................................................................................... min Month ...................................................................................................................................................... mo Newton ...................................................................................................................................................... N Newtons per metre ................................................................................................................................ N/m Ohm (electrical) ......................................................................................................................................... Ω Ounce ...................................................................................................................................................... oz Parts per billion ...................................................................................................................................... ppb Parts per million .................................................................................................................................... ppm Pascal (newtons per square metre) ......................................................................................................... Pa




Pascals per second ............................................................................................................................... Pa/s Percent ...................................................................................................................................................... % Percent moisture (relative humidity) .................................................................................................... % RH Phase (electrical) ..................................................................................................................................... Ph Pound........................................................................................................................................................ lb Power factor ............................................................................................................................................. pF Revolutions per minute .......................................................................................................................... rpm Second (plane angle) .................................................................................................................................. " Second (time) ............................................................................................................................................. s Specific gravity ........................................................................................................................................ SG Square centimetre .................................................................................................................................. cm2 Square kilometre .................................................................................................................................... km2 Square metre ........................................................................................................................................... m2

Thousand pounds ................................................................................................................................... klb Thousand tonnes ...................................................................................................................................... kt Tonne (1,000 kg) ......................................................................................................................................... t Tonnes per day ........................................................................................................................................ t/d Tonnes per hour ....................................................................................................................................... t/h Tonnes per year ....................................................................................................................................... t/a Total dissolved solids ............................................................................................................................ TDS Total suspended solids ......................................................................................................................... TSS Volt .............................................................................................................................................................V Week ........................................................................................................................................................ wk Weight/weight ........................................................................................................................................ w/w Wet metric ton ....................................................................................................................................... wmt




4.0 PROPERTY DESCRIPTION AND LOCATION

4.1 Location

The Mina de Cobre Panamá Concession is located in Colón Province of north central Panamá, approximately 120 km west of Panamá City (Figure 4-1). The process plant site location is N8°50' and W80°38'; the port site location at Punta Rincón is N9°02' and W80°41'.

Figure 4-1: Location Map

4.2 Property Description and Legal Status

Inmet Mining Corporation, directly and through several wholly-owned subsidiaries, owns 100% of the shares of Minera Panamá S.A. (MPSA). Korea Panamá Mining Corp, a wholly-owned subsidiary of LS-Nikko Copper Inc., has an option to earn up to a 20% interest in MPSA by funding its pro rata share of MPSA’s ongoing development costs, to a maximum of US$150 million, until a production decision is made, and re-imbursing its pro rata share of past expenditures.

Under Panamanian Ley Petaquilla, or Law No. 9, 1997, the concession rights to the Mina de Cobre Panamá (then Petaquilla) property were granted to Minera Panamá, S.A. (then Minera Petaquilla, S.A.). This project-specific law gave MPSA rights over the gold, copper, and other mineral deposits for the purposes of exploring, extracting, processing, transporting, and




marketing of all base or precious minerals located in the Concession Area. Later, in June 2005, the Molejón Gold Project Agreement gave Petaquilla Minerals Ltd (PTQ) the property and mineral rights over a portion of the Concession Area, deemed the Molejón Sub-Concession, to permit it to develop the Molejón gold deposit on a stand-alone basis. PTQ was also given the right to explore and mine gold deposits (greater than 50 percent of the present value being derived from gold or other precious metals content) in the larger Concession Area provided such rights do not impair or impead MPSA’s ability or interest to exploit the Concession Area, while MPSA retains the right to develop any copper deposits on the Molejón Sub-Concession. The Molejón Sub-Concession covers 600 hectares of the total 13,600 hectare Concession Area.

MPSA will exercise its rights under Law No. 9, 1997, to acquire or lease state lands located in the proposed tailings basin area. MPSA has undertaken an investigation of existing private holders of surface rights and will initiate negotiations for the acquisition of these properties according to the procedures established by Law No. 9 and other applicable laws. In the event that lands are occupied, MPSA will adhere to International Finance Corporation’s Performance Standard 5 in connection with relocation and resettlement of affected persons and communities. MPSA will also conform with the requirements of IFC PS 6 regarding protected areas, since the proposed TMF is located within the Donoso Multiple-Use Area, a form of protected area in Panamá.

At Punta Rincón, where the proposed port site facilities are located, MPSA purchased surface title to much of the land required for construction and permanent use of the site between 1998 and 2000. Additional land will be acquired as above.

MPSA is entitled under Law No. 9 to an easement for the Coast Road between the Concession and Punta Rincón.

The location of planned project facilities is shown in Figure 4-2.




Figure 4-2: General Arrangement of Project Facilities

The Concession consists of four zones totalling 13,600 hectares as defined in Annex 1 of Law No. 9 -1997 and shown in Figure 4-3. The geographic coordinates for each zone are given in Table 4-1.




Figure 4-3: Property Location Map – Mina de Cobre Panamá Concessions




Table 4-1: MPSA Mineral Concessions under Law No. 9, 1997

Geographic Coordinates – NAD27 UTM Zone 17 (Canal Zone)Zone

Longitude Latitude Direction Distance

(m) Area (ha)

Zona 1 80°41'59.02" 8°51'25.11" East 8,000 4,000 80°37'38.15" 8°51'25.11" South 5,000 80°37'38.15" 8°48'42.07" West 8,000 80°41'59.02" 8°48'42.07" North 5,000 Zona 2 80°45'14.67" 8°54'40.76" East 6,000 6,600 80°41'59.02" 8°54'40.76" South 11,000 80°41'59.02" 8°48'42.07" West 6,000 80°45'14.67" 8°48'42.07" North 11,000 Zona 3 80°40'53.80" 8°48'42.07" East 6,000 1,200 80°37'38.15" 8°48'42.07" South 2,000 80°37'38.15" 8°47'36.85" West 6,000 80°40'53.80" 8°47'36.85" North 2,000 Zona 4 80°37'38.15" 8°50'52.55" East 3,000 1,800 80°36'00.48" 8°50'52.55" South 6,000 80°36'00.48" 8°47'36.85" West 3,000 80°37'38.15" 8°47'36.85" North 6,000 Molejón Sub-Concession 80°39'20.87" 8°48'44.19" East 3,000 600 80°37'42.70" 8°48'44.19" South 2,000 80°37'42.70" 8°47'39.09" West 3,000 80°39'20.87" 8°47'39.09" North 2,000

Zona No. 1, with a total surface area of 4,000 hectares, lies in the Jurisdictions of Northern Coclé and San José del General, Donoso District, Province of Colón.

Zona No. 2, with a total surface area of 6,600 hectares, lies in the Jurisdiction of Northern Coclé, Donoso District, Province of Colón. It is contiguous to the west of Zona No. 1.

Zona No. 3, with a total surface area of 1,200 hectares, lies in the Jurisdiction of San José del General, Donoso District, Province of Colón. It is contiguous to the south of Zona No. 1.

Zona No. 4, with a total surface area of 1,800 hectares, lies in the Jurisdiction of San José del General, Donoso District, Province of Colón. Zona No. 4 is contiguous to the east of Zona 1 and contiguous to the east of Zona 3.

4.3 Permits

In accordance with current Panamanian legislation, MPSA must submit an Estudio de Impacto Ambiental Category III (ESIA, or Environmental and Social Impact Assessment) to the environmental authority, Autoridad Nacional del Ambiente (ANAM, or the National




Environmental Authority). ANAM’s approval of the ESIA will constitute the major permit to proceed with the construction of the project. Submission of the ESIA for approval is planned for April 2010.

More than 150 additional permits have been identified as being required for development, such as the clearing of trees, use of water and construction activities. Other regulatory agencies include the maritime authority for port facilities and the municipality of Donoso for general construction. Application for these permits will proceed in parallel with and immediately following the EsIA approval process.

4.4 Royalties and Taxes

Law No. 9, 1997, which governs the Project Concession, provides for payment of a royalty of 2% of “Negotiable Gross Production” (defined as “the gross amount received from the buyer due to the sale (of concentrates) after deducting all smelting costs, penalties and other deductions, and after deducting all transportation costs and insurance…incurred in their transfer from the mine to the smelter”) as well as payments of 3 Panamanian Balboas (US$3.00) per hectare per year of Concession area.

In addition, Law 9 provides certain fiscal incentives including:

1. “Exoneration for the Company, its Affiliates, contractors and subcontractors of any import tax or duty, contribution, charge, consular fee, lien, duty or another tax or contribution, or of any name or class that fall [are levied] on the introduction and import of equipment, machinery, materials, parts, diesel and Bunker C and other petroleum derivatives”

2. “Income tax exoneration applicable to remittances or transfers abroad, made to pay commissions, loans, royalties, returns, charges for professional advice or administration incurred outside the national territory”

3. “Excepting only the respective mining royalties and royalties, as long as the Company has not finished repaying the debt which the Company or its Affiliates acquire for construction and development of [the Project], the Company and its Affiliates shall be totally exempt from payment of any type of tax, fee, duty, charge, lien, contribution or tribute that may be levied due to any reason in relation to the development of THE PROJECT, except municipal taxes.”

For the purposes for this study it is assumed the term of the debt repayment schedule will be 11 years. These benefits and charges are incorporated into the financial evaluation of the project.

All quotations in this section are from an English translation of Law 9.




5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

5.1 Accessibility

The Mina de Cobre Panamá Concession is located in the district of Donoso, Colón Province in the Republic of Panamá, approximately 120 km west of Panamá City (Figure 4-1). Colón is in the north central part of Panamá, bounded by the Caribbean Sea to the north and the Coclé province to the south. The project will entail two main development areas: a mine and plant site located within the concession boundaries and a port site at Punta Rincón which is located on the Carribbean coast, 25 km north of the plant site. The plant site is located at latitude 8o 50’ North and longitude 80o 38’ West. The port site location is located at latitude 9°02' North and longitude 80°41' West.

Access to the property is via the southern Pan-American Highway from Panamá City to Penonomé, surfaced all weather roads to La Pintada and gravel roads from Coclecito to the Colina camp. There are a number of drill roads on the Botija deposit but access to most of the drill pads at Botija and the other mineralized zones is by helicopter. A large concrete helicopter pad was constructed at the Colina camp and is used for mobilizing crews and drills.

There is an existing airplane runway at Conclecito but the frequent, thick cloud cover in this area is an impediment that would limit the availability of any runway operations.

5.2 Climate

Climatic conditions are equatorial. Annual precipitation averages 4,700 mm, humidity is high, and temperatures are relatively high (25°C to 30°C) year-round. Coastal areas generally receive more rainfall than inland locations. Heavy tropical rains are prevalent throughout the year – even the driest month receives more than 60 mm of rain. Storms are generally of short duration, ranging from 1½ to 2 hours.

5.3 Local Resources and Infrastructure

There is no industrial development in this part of Panamá, and the region is sparsely populated. Subsistence farming is the primary occupation of the local residents. The nearest community to the project site is the village of Coclecito, 12 km southeast of the proposed plant site. The city of Penonomé, the capital of Coclé province, is 49 km further southeast of Coclecito.

MPSA has two main camps on the property and several smaller ones in more remote parts of the property. The Colina camp and New camp can accommodate approximately 400 people and both are accessible by road. Core shacks, core storage and sample preparation facilities are present at both camps. Remote camps are located at the Colina and Botija deposits.

The development of the Cobre Panamá Project will require the construction of a port facility and 300 MW power plant at Punto Rincón, about 25 km north of the mine and plant area. A Coast Road must be constructed to connect these two sites. Other required development includes




power transmission lines, concentrate and water pipelines, man camps, shops, and other support facilities. The high rainfall environment will provide ample water for the project, which will be captured in the tailings basin and other collection areas.

The Botija, Colina and Valle Grande mineral deposits are within the existing concession boundaries. MPSA has rights to acquire or lease lands for waste rock storage and tailings facilities, stockpiles, mill and port sites, and other support facilities.

While local people will be given hiring preference, additional personnel will be required from other areas of the country. Substantial training programs will be required for national workers. Skilled expatriate personnel will be needed in the early stages of the project to help provide this training and for initial project management.

5.4 Physiography

The Concession area is characterized by rugged topography with heavy rainforest cover. The dominant landforms are relatively narrow ridges that parallel major geological structural trends and are bisected by numerous surface water drainages. Elevations range between 70 m asl and 300 m asl in the vicinity of the Botija, Colina and Valle Grande deposits. The terrain at the port site is gentler, with elevations ranging from sea level to approximately 60 m asl.




6.0 HISTORY

Copper-gold-molybdenum porphyry mineralization in the Río Petaquilla region of central Panamá was discovered in 1968 by a United Nations Development Program team who were conducting regional geological and geochemical surveys. Exploration by several companies has since outlined three large porphyry deposits, a host of smaller mineralized zones, and one zone of epithermal gold mineralization.

Several pre-feasibility and feasibility studies have been done. These include:

1. A preliminary feasibility report was prepared by Panamá Mineral Resources Development Co. Ltd. in 1977 and updated in 1979

2. Kilborn Engineering Pacific Ltd completed a prefeasibility study in 1994 and an update in 1995 for Adrian Resources Ltd., a precursor of Petaquilla Minerals Ltd. and Petaquilla Copper Ltd.

3. H.A. Simons, now AMEC, produced a feasibility study for Teck Corporation in November 1996 with a subsequent update in January 1998. The resource estimates included in the 1998 are discussed in more detail in section 17.1.14.

In May 1998, the updated January 1998 document was submitted to Dirección General de Recursos Minerales (General Directorate of Mineral Resources, DGRM) of the Panamanian Ministry of Industry and Commerce, and was accepted as the official Feasibility Study to satisfy concession law requirements for the delivery of a feasibility study. The Petaquilla (Mina de Cobre Panamá) concession rights were granted to Minera Petaquilla, S.A., now Minera Panamá, S.A. (MPSA), under Panamanian Law No. 9 on 26 February 1997. At that time, the shareholders of MPSA were Petaquilla Copper Ltd., Teck Cominco, and Inmet Mining Corporation.

In June 2005, through the Molejón Gold Project Agreement, MPSA transferred to Petaquilla Minerals Ltd. the property and precious-metal mineral rights over a portion of the Concession Area, deemed the Molejón Sub-Concession, to permit it to develop the Molejón gold deposit on a stand-alone basis.

In September 2008 Inmet acquired Petaquilla Copper Ltd, and in November 2008 Inmet acquired Teck Cominco’s remaining share in MPSA, taking Inmet to a 100% interest in MPSA.

In October 2009 Inmet announced an option agreement with LS-Nikko Copper Inc. under which LS-Nikko has the right to acquire a 20% interest in the Mina de Cobre Panamá copper project. If LS-Nikko exercises the option, it will receive an equity interest in MPSA.

The name of the project was changed to Mina de Cobre Panamá in 2009.

The exploration history of the Cobre Panamá Concession is summarized briefly in Table 6-1.




Table 6-1: Exploration History of the Cobre Panamá Concession

Year Party Description

1968 UNDP Regional geological and geochemical survey of central Panamá by the United Nations Development Program (UNDP); widespread silicification and copper mineralization discovered in area of Colina and Botija deposits; silt samples and 200 line-km of soil samples revealed several copper and molybdenum anomalies, including Valle Grande, Botija Abajo, Brazo, and Medio; vertical field magnetics identified areas of magnetite alteration and magnetite destruction.

1969 UNDP 27 short (Winkie drill) holes and 10 long holes drilled in Botija, Colina, Vega, and Medio areas.

1969 PMRD Panamanian government tendered Cobre Concession exploration rights to international bidding; concession awarded to Panamá Mineral Resources Development Company (PMRD), a Japanese consortium.

1970-1976

PMRD Geological mapping at Botija and Colina; 48 short (Winkie drill) and 51 long (diamond) holes drilled at Botija, Colina, Medio, and Vega, totalling approximately 14,000 m. Botija and Colina deposits drilled on approximately 200 m centres.

1977 PMRD Preliminary reserves calculated and pre-feasibility report completed.

1978-1979

PMRD Feasibility work updated; unsuccessful negotiations with the Panamanian government over terms of production.

1980 PMRD Property abandoned by PMRD.

1990-1992

Minnova Property acquired by Minnova (now Inmet), 80%, and Georecursos International S.A., 20%. Exploration activity included regional lithogeochemical sampling.

1992-1993

Adrian Adrian Resources Ltd. (Adrian) granted an option to earn 40% of Minnova’s interest through cash payments, work commitment, and production of a feasibility study. Adrian subsequently acquired Georecurso’s interest, bringing its total interest to 52%.

1992-1995

Adrian Adrian carried out grid-based soil sampling and magnetic measurements, geologic mapping of selected areas, and drilling of approximately 396 diamond drill holes in Colina, Botija, and exploration targets. Investigation of Valle Grande deposit and discovery of epithermal Au mineralization at Molejón, as well as identification or investigation of several other targets (Botija Abajo, Brazo, Faldalito, Cuatro Crestas, Lata, Orca). Initiation of baseline environmental studies. Scoping study and pre-feasibility study produced.

1994 Teck Teck was granted the right to acquire half of Adrian’s share (26%) of the deposit by funding a feasibility study and arranging Adrian’s portion of the financing needed to bring the deposit into production.

1996 Teck Infill and deposit condemnation drilling and mapping for feasibility study carried out. Teck drilled 91 infill and 33 condemnation holes totalling 26,837 m. Feasibility study completed.




Year Party Description

1997 Teck Infill and drilling for metallurgical samples to update feasibility study. Teck drilled 43 holes totalling 8,099 m. Feasibility study updated.

2005 MPSA Molejón Gold Agreement – shareholders transfer rights to any gold deposits on concession to Petaquilla Minerals Limited for a 5% NSR.

2007-2008

MPSA Activity resumes on copper deposits with the JV drilling condemnation, metallurgical, infill, and pit geotech holes Lidar topo survey of concession.

2008 Inmet Inmet aquires Petaquilla Copper Ltd (PTC) including the 26% interest in MPSA, taking Inmet to 74% interest in MPSA.

2008 Inmet Inmet acquires Teck’s 26% interest in MPSA, taking Inmet to 100% interest in MPSA.

2009 Inmet 2007-2009: 288 holes (73,481 m) of infill, metallurgical, and condemnation drilling at Botija, Colina, Valle Grande, and Brazo deposits. Comminution testing. Condemnation of proposed tailings area and seismic, resistivity, and geotech drilling at port site and infrastructure locations.

2009 KORES/LS Nikko Copper

KORES and LS Nikko agree to option a 20% interest in MPSA.




7.0 GEOLOGICAL SETTING

7.1 Regional Geological Setting

Panamá is a tectonically active area at the junction of four lithospheric plates. South vergent subduction and related arc volcanism, high-angle strike-slip and block faulting, and north vergent thrust faulting are currently shaping the country (Mann, 1995). Western Panamá and eastern Costa Rica are underlain by the Chorotega crustal block and share similar characteristics. This block is underlain by a late Cretaceous to recent volcanic arc, constructed on a basement of late Cretaceous to Palaeocene oceanic crust and marine sedimentary and volcanic rocks (Figure 7-1). Despite limited exploration and geological mapping, a number of significant copper porphyry deposits have been discovered, including the Cobre Panamá deposits, Cerro Colorado (1.52 billion tonnes averaging 0.78% Cu; Sides, 1994) and Cerro Chorcha (117 million tonnes averaging 0.51% Cu; Bellhaven Resources website).

Figure 7-1: Generalized Geology of Panamá




In the Chorotega block, the island arc sequence consists of several distinct pulses of volcanism, including Palaeocene-Eocene, mid-Oligocene, late Oligocene to early Miocene, and Pliocene-Pleistocene, probably separated by times of plate reorganization (de Boer et al., 1995).

Intrusive rocks of the older suite lie along a tholeiitic trend on an AFM plot. No porphyry mineralization is recognized during this period (Kesler et al., 1977). Younger rocks are calc-alkaline and contain porphyry mineralization ranging from Oligocene (Mina de Cobre Panamá deposits 32 Ma) to Pliocene (Cerro Colorado 5 Ma).

Middle Oligocene rocks in the Chorotega block, including the 400 km2 Petaquilla batholith, range from gabbros to hornblende granites. The more northerly location of the Petaquilla batholith relative to the axis of the older arc suggests a flattening of subduction (de Boer et al., 1995). Plutonic rocks of Miocene and younger age are progressively more felsic and calc-alkaline, with an increase in K2O, corresponding to the evolution of the volcanic arc over time (Kesler et al., 1977). DeBoer et al. (1995) note that Oligocene rocks typically exhibit negative zirconium anomalies, whereas the Miocene and younger rocks show positive zirconium anomalies. Despite these trends, intrusive rocks of all ages exhibit low 87Sr/86Sr initial ratios (Kesler et al., 1977), suggesting derivation from mantle wedge and slab, and emplacement into relatively primitive oceanic basement.

In the area of the Cobre Panamá deposits, the oldest rocks are submarine andesite and basalt flows and tuffs, intercalated with clastic sedimentary rocks and reef limestones, of probable Eocene to early Oligocene age. The arc became emergent during mid-Oligocene time, with terrestrial flows and volcaniclastic rocks and lesser intercalated submarine tuffs. Miocene and younger rocks comprise the bulk of volcanic rocks in western Panamá and consist of both terrestrial and marine volcanic and volcanic-derived rocks of progressively more felsic composition.

Reconnaissance mapping (Figure 7-2) in the region of the MPSA Concession and surrounding areas in central Panamá (Azuero Peninsula) was undertaken from 1966 to 1969 by the United Nations Development Program (UNDP, 1968, 1969; Kent, 1968). The region is underlain by altered andesitic to basaltic flows and tuffs and clastic sedimentary rocks of presumed early to mid-Tertiary age, intruded by the mid-Oligocene (36.4 ± 2 Ma hornblende K-Ar age; Kesler, 1977) Petaquilla batholith, which is of granodiorite composition. Numerous satellite stocks and plutons composed of equigranular to porphyritic granodiorite, tonalite, quartz diorite, and diorite occur around the margin of the batholith, and are most prevalent along its southern boundary. Interpretation of satellite imagery for the region suggests that major structural trends, expressed as topographic linears, are northeast and northwest (Figure 7-2). Northwest-trending linears are parallel and may be related to the Canal shear zone and other large left-lateral shear zones in the region.




Figure 7-2: District Geology and Structural Lineaments

A more detailed geological map for the concession is based on work done Adrian Resources in the 1992’. The results of this work is shown in Figure 7-3.




Figure 7-3: Property Geology

Note Adapted from 1992 – H.J. Awmack, Adrian Resources

7.2 Lithology

Seven main lithological units have been recognized on the property. From oldest to youngest these units are:

1. a sequence of undifferentiated andesite volcanics and volcaniclastics

2. granodiorite

3. a suite of a feldspar quartz hornblende porphyries and that are likely to be of granodiorite composition

4. a set of feldspar quartz porphyry dikes

5. a suite of fine-grained mafic dikes

6. a suite of andesite porphyry dikes

7. saprolite.

Cross-cutting relationships suggest that the andesite volcanics and volcanoclastics were intruded by the granodiorite, which forms a batholith and hosts most of the copper




mineralization. This was followed by the intrusion of dikes and apophyses of the feldspar-quartz-hornblende porphyry that is the most extensive of the porphyry lithologies and contains slightly higher grade copper mineralization than the granodiorite. There are few cross-cutting contacts between the feldspar quartz hornblende porphyry and the granodiorite, and at times it is difficult to texturally distinguish the two lithologies. This suggests that they were intruded over a short period of geological time, most likely as a continuum of differentiation products of the regional-scale batholith. The intrusion of a volumetrically minor set of feldspar quartz porphyry dikes followed, which so far has only been recognized at Botija and Brazo.

A set of mafic dikes cuts the aforementioned lithologies; the dikes are volumetrically very minor and may be a final differentiation phase of the batholiths. A set of late andesite dikes clearly cross-cuts all other lithologies and exhibits well-developed chill margins, indicating that the dikes were intruded after the main porphyry intrusive event had cooled. Since exhumation of the porphyry system, surface weathering has produced a saprolite profile that is typical of tropical environments.

The lopolith morphology of the feldspar quartz hornblende porphyry bodies at Botija, Colina, and Valle Grande have led to consideration of magmatic stoping as its likely mechanism of emplacement. Evidence for stoping includes:

the abundance of roof pendants

the often flat-lying geometry of the lower porphyry contact

the proximity of a large batholith and likely underlying magma chamber, the deflation of which would be capable of allowing subsidence of the lower contact

the presence of fracturing at the lower porphyry contact without significant displacement – a feature typical of the dilatational opening required by such a model.

Space for emplacement of the feldspar quartz hornblende porphyry in the three deposits was likely created by cantilever subsidence of the lower contact of the intruding body (Cruden, 1998). At Botija and Colina the feldspar quartz hornblende porphyry is interpreted to have flowed laterally into a sill-like lopolith that was fed from a high-angle feeder dike in the north of the deposit. At Valle Grande emplacement was likely fed from a high-angle feeder dike to the northwest. Sinking blocks are only “trapped” in the final crystallization phase of a magma (Žák et al, 2006). This is consistent with the interpretation that the feldspar quartz hornblende porphyry in the three deposits formed as a differentiation product of the larger granodiorite batholith.

A detailed description of each of these lithologies follows, with the respective logging code provided in parentheses.

Andesite (ANDS) – Light green, chlorite-altered, fine-grained andesite that is weakly magnetic and comprises mostly massive flow units that are bounded by fragmental flows. Lesser lapilli tuffs and volcanoclastic interbeds have also been noted. The andesites have not been subdivided into individual flows in the current logging scheme.




Granodiorite (GRDR) – Mottled white-grey lithology with medium-grained equigranular interlocking crystals of white feldspar hornblende and quartz. The granodiorite also contains subrounded-subangular, centimetric fine-grained mafic xenoliths and occasional aplite dikes. White clay alteration can produce a porphyritic texture by altering the groundmass surrounding feldspar crystals, producing a pseudo-porphyritic texture and sometimes causing it to be logged as the Feldspar Quartz Hornblende Porphyry.

Feldspar Quartz Hornblende Porphyry (FQHP) – Light grey porphyritic lithology with a crowded porphyry texture. There is not always a large difference in the grain size of the groundmass, and the phenocrysts. The porphyritic texture is best described as weak. The groundmass is light grey, fine-grained, and contains feldspar with subordinate mafics and quartz. Phenocrysts comprise crowded white subhedral to euhedral feldspar (often plagioclase) phenocrysts, lesser clear glassy anhedral quartz “eyes,” and minor hornblende.

Feldspar Quartz Porphyry (FQP) – Light grey porphyritic lithology with a light grey, aphanitic and siliceous groundmass that contains crowded white subhedral to anhedral feldspar phenocrysts and clear, glassy anhedral quartz “eyes” and very rarely bi-pyramidal quartz phenocrysts. This lithology has so far been recognized at the Botija and Brazo deposits only.

Feldspar Hornblende Porphyry (FHP) – Light green-grey porphyritic lithology, probably of andesitic composition, with a light grey, fine-grained groundmass containing dominant white subhedral to euhedral feldspar (mostly plagioclase) with subordinate dark green hornblende and lesser quartz. This porphyry contains sparse, irregularly shaped, centimetric, medium-grained mafic xenoliths.

Mafic Dikes (MD) – Dark green, pervasively chlorite altered, fine-grained mafic pyroxene porphyry that occurs as dikes with clear chill margins noted locally.

Saprolite (USAP, LSAP, SPRC) – At surface, most areas of the property are covered by a thin layer (1 cm to 10 m) of black organic material. This overlies a thin layer of residual soil generally less than 1 m in thickness, which in turn overlies a 1 m to 20 m thick layer of orange to white (oxidized) or green (reduced) saprolite. On average, the saprolite layer is deeper over andesite relative to intrusive rocks. The transition from saprolite to relatively fresh andesite typically takes place over a range of less than a few metres, whereas the transition zone from saprolite to relatively fresh granodiorite may involve several tens of metres. This transition zone contains blocks of unaltered bedrock and has been called “saprock.”

7.3 Alteration

Five types of alteration were modelled at Botija, Colina, and Valle Grande. Propylitic alteration was divided into types A and B. The alteration types are listed below, beginning with the earlier one and progressing to assemblages that likely overlap or occur later in the development of the hydrothermal system.




Propylitic A alteration – Chlorite dominated with accessory epidote, pyrite, and calcite. It is particulary well developed within the andesite volcanics.

Propylitic B (silica-chlorite) alteration – Exhibits chlorite and silica in approximately equal intensity, resulting in a much harder core that other alteration types. In porphyry lithologies chlorite generally affects ferro-magnesium phenocrysts while silicification appears to primarily affect the groundmass. Pyrite and sericite (often green-coloured) may also be present. Previous workers on the property have referred to this alteration type as silica-chlorite alteration. It appears to occur at depth within the deposits but is found at shallow levels in the peripheral zones.

Potassic alteration – Occurs mainly as potassium feldspar selvedges to quartz ± sulphide veinlets and as irregular patches. Potassium feldspar flooding is rarely seen. Potassic alteration also occurs as fine-grained secondary biotite that alters ferro-magnesium minerals such as hornblende and magmatic biotite. Secondary biotite also occurs in discontinuous veinlets that commonly contain magnetite, chalcopyrite, and rare bornite. The amount of potassium feldspar or secondary biotite alteration is largely determined by the feldspar or biotite abundance in the protolith. Anhydrite veinlets of generally millimetric thickness are commonly associated with potassic alteration, especially in the deeper parts of the deposit. At depths of approximately 200 m from surface and shallower, the anhydrite appears to have hydrated to form gypsum. In addition, millimetric magnetite-only veinlets are uncommonly observed within the potassic alteration zone.

Phyllic alteration – Occurs as sericite alteration of all rock-forming silicate minerals. Silicification and pyrite are also associated with this phase of alteration. Phyllic alteration occurs in the upper 150 m to 200 m of all of the deposits but is very irregular and/or difficult to map through different protoliths and earlier alteration facies. It can occur much deeper when phyllic fluids have been drawn down into permeable structural zones. Phyllic alteration occasionally occurs with chlorite, and both green and white sericite are found within this zone, likely related to distinct alteration events. Sericite is used as a collective term for white phyllosilicate and displays a wide range of grain size from very coarse, granular, millimetric muscovite, to very fine grained “silky” textured. It may be possible to subdivide these variations spatially because they are probably related to distinct alteration events. Phyllic alteration is ubiquitous, with quartz-pyrite veinlets that frequently contain minor chalcopyrite and have a white sericite-altered selvedge.

Argillic alteration – Frequently occurs in the upper parts of the deposits from surface and is therefore largely coincident with the phyllic alteration zone. This can make visual distinction of clay or sericite dominant alteration a challenge. The clay minerals that have been visually identified range in colour from white to buff or light brown; kaolinite, smectite, and illite have been recognized.

Paragenetic, cross-cutting, and mineral texture relationships suggest that the two propylitic mineral assemblages and the potassic alteration formed early, and both are frequently associated with chalcopyrite, minor bornite, and minor pyrite mineralization. These early




alteration types are overprinted by phyllic alteration that includes white and green sericite and ubiquitous pyrite with variable silicification or quartz veining. In addition to pyrite, phyllic alteration has been observed to occur frequently with chalcopyrite, but very rarely bornite. Argillic alteration overprints propylitic, potassic, and phyllic alteration and is therefore paragenetically younger. In general argillic alteration occurs within 300 m from surface and is most common near surface. White clay found within a few tens of metres of oxidized sulphides is most likely supergene in origin, perhaps overprinting an earlier alteration phase. Other clay minerals can be assumed to relate to hypogene alteration. Argillic alteration has been modelled from visual observation of clay. Little resolution on clay mineralogy is available, and there may be some overlap with logged sericite.

In addition to the five dominant alteration types described above, metasomatism of the andesite volcanic package can produce a biotite hornfels with significant disseminated sulphide mineralization that is likely broadly coeval with the formation of propylitic and potassic alteration.

Post-mineral zeolite alteration in the intrusion and andesite is widespread and consists of pink zeolite veinlets and fracture fills with vein-hosted calcite and rare barite, locally making up as much as 10% of the rock mass. This likely occurred sometime late in the development of the phyllic alteration and before hypogene clay alteration ended. Chloritization of mafic minerals may also accompany this late-stage mineral assemblage.

7.4 Botija

Drill roads have provided some outcrop exposures at Botija but most of the geological knowledge has been a result of drilling. A plan map of the surface geology was constructed based on surface mapping that has been supplemented with the collar geology for each drill hole (Figure 7-4).

The granodiorite and the feldspar quartz hornblende porphyry host two irregular, keel-shaped andesite roof pendants measuring approximately 500 m in diameter and separated by approximately 300 m. The roof pendant in the central part of the deposit reaches a depth of 200 m. The other andesite roof pendant is located at the eastern margin of the deposit and extends to a depth of at least 300 m. A smaller north-south elongate andesite roof pendant measuring approximately 250 m x 100 m and extending to a depth of 150 m sits between the northern margins of the two larger bodies.




Figure 7-4: Botija Deposit - Plan Map - Geology, Drill Hole Locations

Note: Andesite (green), Porphyry (orange), Granodiorite (pink)

A set of north-south geological cross sections spaced at 50 or 100 meter intervals has been constructed over the Botija mineralized zone. A geological model illustrating the distribution of the main rock types was constructed using these sections. An example of one of these cross sections is given in (Figure 7-5).

The feldspar quartz hornblende porphyry appears to have intruded as one to four dikes ranging in thickness from 20 m to 200 m that coalesce to form a single body between 100 m and 600 m thick that extends down to a depth of approximately 450 m. In a general sense, mineralization shows a spatial association with this intrusive phase. The feldspar quartz hornblende porphyry morphology gives the impression that it intruded from the north as a series of dikes that fed the central apophysis, where they coalesced. In general the dip of the more distinct dikes is approximately 70° to the north. The feldspar quartz hornblende porphyry appears to support the andesite roof pendants rather than cross-cut them. This feature may indicate that the porphyry was emplaced relatively passively by magmatic stoping rather than as a stock that could be expected to cross-cut pre-existing lithologies. There is very little hydrothermal or magmatic brecciation at Botija.




Figure 7-5: Section 53 8200 East - Botija Deposit


At Botija, propylitic alteration is irregular and not dominant, occurring generally sporadically at depth and in the periphery of the deposits. Propylitic A alteration is associated with andesite specifically. Potassic alteration is widespread in the central part of the deposit and shows a loose spatial association with higher copper grades and porphyry intrusive, often following the northerly dip of these features. Phyllic alteration is irregular, occurring mainly in the central area of the deposit near surface and also at depth. Argillic alteration is also irregular and mostly occurs within 250 m of surface.

At Botija, displacement of lithological units and mineralization was used in an attempt to identify the location of significant faults, and three minor faults were traced from surface through the drill sections using the available data. However, displacement of lithology was a few tens of metres across these faults at the most, and mineralization and alteration were not significantly displaced.

Zones of increased fracturing were noted on the southern margin on Botija, along the previously described Botija River Fault Zone (Speidel and Faure, 1996), but evidence of displacement was not found, and insufficient data exist to confidently model it in three dimensions.

Numerous centimetric high-angle faults with some gouge have been noted in drill core for Botija, but they are too numerous to connect in a coherent model. These minor faults trend 070°-090° and 300°-330° and commonly contain pyrite and sometimes chalcopyrite, frequently with chloritic and sericitic feldspar quartz porphyry dikes in the western part of the deposit




(Escalante, 2009). Most of these faults dip to the north. Another easterly dipping centimetric fault set with minor displacement trends 160°-180°, dips between 20° and 60°, and may control the distribution of chloritic, sericitic, and clay-altered feldspar quartz porphyry dikes and local quartz veining in the north and southern part of the deposit (Escalante, 2009). One of these faults may also constitute the southern contact of the eastern andesite roof pendant at Botija (Escalante, 2009). If stoping is considered to be the dominant mechanism for emplacement of the feldspar quartz hornblende porphyry at Botija, then many of the zones of minor faulting and fracturing could be caused by dilation, and minor dip or strike-slip faults could be expected, particularly in the subsiding floor of the feldspar quartz hornblende lopolith.

7.5 Colina

Outcrop exposures in the Colina area are found in creek bottoms and at recently constructed (i.e. not overgrown) drill pads. A plan map of the surface geology was constructed using the collar geology for each drill hole (Figure 7-6).

Figure 7-6: Colina Deposit - Plan Map - Geology, Drill Hole Locations


The Colina deposit is focused on a 2.5 km long x 1 km wide feldspar quartz hornblende porphyry sill and dike complex (lopolith) that trends east-southeast and has been closed by drilling in this direction; remaining open to the west-northwest. Located in the southern part of the deposit is an irregular L-shaped andesite roof pendant with the long axis orientated to the




east-southeast and the short axis orientated to the south-southwest. The long axis is 100 m to 200 m wide at surface, the short axis 400 m to 500 m wide.

A set of north-south geological cross sections spaced at 100 meter intervals has been constructed over the Colina mineralized zone. A geological model illustrating the distribution of the main rock types was constructed using these sections. An example of one of these cross sections is given in (Figure 7-7).

The majority of the feldspar quartz hornblende porphyry comprises 50 m to 200 m thick sills that dip shallowly to the north and are often interconnected by dikes. The sills and dikes coalesce in the centre of the deposit, offset towards the northern contact from where the porphyry appears to emanate.

Figure 7-7: Section 53 3800 East - Colina Deposit


At the base of the feldspar quartz hornblende sills, a transition to granodiorite often occurs, accompanied by increased fracturing, frequently with the presence of some minor mafic dikes, and, in a general sense, frequently with a rapid decline in copper grade. These contacts are thought to correspond to the lower stoped contact and the fracturing to be largely due to tensional forces exerted by the sinking granodiorite causing tensional failure. Ingress of highly fractionated residual mafic magma locally may explain the spatial relationship to narrow mafic dikes in these areas. At the northern contact of the dike, the granodiorite and andesite host rocks are intermixed, and the nature of the contact and morphology of the granodiorite are complex. In general, the emplacement mechanism is thought to correspond to a lower contact




slab, “hinged” in the south with a cantilevered down-dropping edge to the north, where sills and feeder dikes of feldspar hornblende quartz porphyry were emplaced.

Propylitic alteration generally affects the andesite in the periphery of the Colina deposit and the andesite in the central part is frequently affected by silica-chlorite alteration. Phyllic alteration is patchy and difficult to interpret as a continuous zone. In general, it is not associated with the higher-grade part of the deposit.

There is very little potassic alteration at Colina compared to Botija, possibly because the feldspar quartz porphyry forms a series of sills with interconnecting dikes rather than a more vertically continuous body. Anhydrite and its (supergene) alteration product, gypsum, are also not as common as they are at Botija, probably because these minerals generally accompany potassic alteration. Magnetite is more common at Colina than Botija in the western and northern parts of the deposit, and an area of quartz-magnetite veining follows one of the thicker sills. Where it is present, potassic alteration occurs as weak potassium feldspar with patchy biotite alteration of mafic minerals.

Weak phyllic alteration is common, and strong sericite alteration is discontinuous but generally found to mantle the elongate core zone of higher-grade copper mineralization

Argillic alteration forms as an often-continuous zone from surface to depths ranging from 20 m to 80 m. It is most likely of supergene origin.

Late, mostly fracture-filling zeolite and calcite alteration is widespread at Colina.

No faults have been modelled at Colina.

7.6 Valle Grande

Outcrop exposures in the Valle Grande area are found in creek bottoms and at recently constructed (i.e. not overgrown) drill pads. A plan map of the surface geology was constructed using the collar geology for each drill hole (Figure 7-8).

The Valle Grande deposit is focused on a 2,000 m long x 500 m wide irregular feldspar-quartz-hornblende porphyry lopolith that trends to the southeast. An irregular andesite roof pendant approximately 300 m in diameter is located at the southeastern end of this body. The lopolith remains open to the northwest.




Figure 7-8: Valle Grande Deposit - Plan Map - Geology, Drill Hole Locations


A set of geological cross sections oriented at 30o east of north and spaced at 100 meter intervals has been constructed over the Colina mineralized zone. A geological model illustrating the distribution of the main rock types was constructed using these sections. An example of one of these cross sections is given in (Figure 7-9).

The northeastern contact of the lopolith is interpreted to be irregular but is generally vertical. The southwestern contact dips to the northeast, meaning that the body widens to surface. In detail, along the southwestern contact the lopolith appears to branch into various 10 m to 200 m wide discontinuous sills that intrude the andesite (with lesser granodiorite) host rock at low angles that reach horizontal locally. It is thought that this area constitutes the feeder dike for the lopolith.




Figure 7-9: Section 13 West - Valle Grande Deposit


At Valle Grande the most common logged alteration type is propylitic. This is likely due to the dominance of andesites as host rocks. The preponderance of this alteration type within the feldspar quartz hornblende porphyry may be due to some wallrock assimilation providing for a higher proportion of ferro-magnesium minerals.

Silica-chlorite alteration is irregular, often occurring deep within the feldspar quartz hornblende porphyry dike.

Potassic alteration is generally erratic in distribution and often occurs at depths of at least 200 m. The thickest, most coherent areas of potassic alteration are in the central part of the principal dike.

Phyllic alteration is irregular and occurs generally on the contacts of the feldspar quartz hornblende dike, clearly mantling it in some drill sections.

Argillic alteration occurs from surface to depths of 100 m to 150 m, often as a continuous zone. There are sporadic areas of deep (200 m to 300 m) argillic alteration that do not correlate well between 100 m spaced sections. Clay that begins at surface is likely to be at least partly supergene, and the isolated, deep patches of clay alteration may be hypogene in origin.

An attempt was made to interpret structural data for Valle Grande for this study, but it has not been possible to identify any specific structural features.




8.0 DEPOSIT TYPES

The mineralized zones on the Cobre Panamá property are examples of Cu-Au-Mo porphyry deposits (Guilbert and Lowell, 1974; Lowell and Guilbert, 1970). Common features of a porphyry deposit include the following:

Large zones of hydrothermally altered rocks that commonly grade from a central potassic core to peripheral phyllic, argillic and propylitic altered zones.

Mineralization is generally low grade and consists of disseminated, fracture, veinlet and quartz stockwork controlled sulphide mineralization.

Mineralization is commonly zoned with a chalcopyrite-bornite core and peripheral chalcopyrite-molybdenite and pyrite.




9.0 MINERALIZATION

9.1 Introduction

Hypogene Mineralization

Hypogene mineralization within the granodiorite and porphyry phases consists of disseminations, micro-veinlets, fracture fillings, veinlets, and quartz-sulphide stockworks. Copper mineralization occurs as chalcopyrite with lesser bornite. Within all of the deposits the proportion of bornite relative to chalcopyrite shows a loosely defined increase with depth. Molybdenite is present, particularly in quartz “B” veinlets (Gustafson and Hunt, 1975). Pyrite is ubiquitous, but the tenor increases in areas of phyllic and chlorite-silica alteration compared to other alteration zones, particularly the potassic. Within the phyllic zone, pyrite occurs in disseminations and in “D” veinlets (Gustafson and Hunt, 1975) with quartz. Specularite and magnetite mineralization also occur as disseminations and veinlets in all of the deposits.

At the andesite contact with the feldspar quartz hornblende, porphyry copper mineralization can reach a high tenor in zones of biotite hornfels. In such areas chalcopyrite is the dominant sulphide, with minor pyrite and rare bornite in veinlets, blebs, and disseminations. This contact-style of mineralization is typically cross-cut by quartz-sulphide veinlets.

The location of zones with hypogene mineralization is shown in Figure 9-1.




Figure 9-1: Location Map of Mineralized Zones

Supergene Mineralization

Recent oxidation of sulphides near surface has leached copper from the present-day saprolite and previously eroded surficial oxidized layer. This has been weakly and irregularly reprecipitated in the upper reaches of the sulphide facies. Secondary sulphides are dominantly chalcocite with minor covellite and rare native copper and occur as fracture fillings, coatings on sulphide minerals, and disseminations. Where these sulphides have been oxidized, malachite is the main copper oxide mineral.

The absence of a significant zone of enrichment at Botija, Colina, and Valle Grande is likely due to the removal by erosion of a well-developed phyllic alteration zone that may have overlain these deposits. The oxidation of pyrite in such a zone supplies sufficient acid for leaching of copper. Additionally, the neutralized host rock of the phyllic zone provides a suitable host rock within which secondary copper minerals can be re-precipitated. A well-developed phyllic alteration zone does exist at Brazo, and significant secondary copper sulphide mineralization is found there.




9.2 Botija

A quartz vein stockwork of “A” and “B” veinlets occurs within the central feldspar quartz hornblende porphyry dike complex and in its contact zones particularly, and appears to form a locus for higher-grade copper and molybdenum mineralization (Gustafson and Hunt, 1975). This stockwork hosts most of chalcopyrite and bornite mineralization, with the latter vein type containing the largest proportion of molybdenite. In the southeast and northern parts of the deposit, narrow zones of epidote and chlorite skarn measuring up to a few tens of metres occur at the andesite contact with specularite, chalcopyrite, pyrite, and often calcite, quartz, bornite, and magnetite veinlets, blebs, and disseminations. Secondary chalcocite and covellite have been observed near the upper contact of the sulphide zone.

9.3 Colina

In general, higher copper grades appear to be loosely spatially associated with feldspar quartz hornblende porphyry, particularly in the thicker areas where dikes and sills coalesce and around the upper contacts of sills. Colina appears to consist of several thin, stacked sheets of high-grade mineralization, separated by areas of low-grade mineralization, and the “sheets” correspond spatially to the aforementioned sills. Pyrite and chalcopyrite with lesser bornite occur mainly in quartz veinlets, frequently with molybdenite. Vein types have not been systematically mapped at Colina, but it is likely that they emanate from the contact areas of feldspar quartz hornblende porphyry sills and dikes that have a complex morphology. Contact metamorphism forms strongly silicified or biotite-altered zones where andesite is in contact with feldspar quartz hornblende porphyry. Pyrite and chalcopyrite are generally found in veinlets with lesser blebs and disseminations. Magnetite and molybdenite have also been noted in these zones, the latter occurring in quartz “B” type veinlets (Gustafson and Hunt, 1975).

Secondary chalcocite and covellite have been observed as sooty coatings, primarily on chalcopyrite located at the base of the saprolite or within partially oxidized structures that penetrate more deeply into the underlying sulphide domain, following permeable structures. Occasionally chalcocite has completely replaced the pre-existing sulphide to form veins and disseminations; native copper rarely occurs in fractures. Locally, within the saprolite or secondary copper zones, malachite is observed where a copper sulphide mineral has oxidized. Secondary chalcocite and covellite have been observed near the upper contact of the sulphide zone.

9.4 Valle Grande

In general, higher-grade copper mineralization occurs along the flanks of the lopolith, following the feldspar quartz hornblende porphyry contacts. In these areas quartz-sulphide veinlet densities are high. Conversely, copper grades and quartz-sulphide veinlet densities are lower within the central axis of the dike. This area corresponds to an elongate, low-grade core of the copper porphyry system that is characteristic of this deposit type.




9.5 Brazo

The Brazo zone is 3 km south-southeast of the Botija deposit. Soil surveys outlined a well-defined Mo anomaly associated with this mineralization. Drilling by Adrian Resources discovered porphyry-style Cu-Au mineralization, which occurs in feldspar-quartz porphyry. Alteration consists of pervasive sericite and/or clay with pyrite and quartz. Well-developed quartz stockworks are present. Supergene mineralization consisting of chalcocite and rare native copper extends locally to a depth of 150 m. Hypogene mineralization consists of chalcopyrite, pyrite, and rare bornite. On the basis of previous drilling, the Brazo zone has a southeastern trend with dimensions of at least 300 m long by 200 m wide and extends to a depth of 350 m. The zone is open to the northwest and southeast. MPSA is currently drilling this zone to assess its size and grade.

9.6 Botija Abajo

The Botija Abajo area is 2 to 3 km southeast and along strike from the Botija deposit and 1 to 2 km northeast of the Brazo zone. Previous drilling in the area by Adrian Resources tested well-defined Au-Mo-Cu soil anomalies. Gold-enriched porphyry copper mineralization is hosted in feldspar-quartz-hornblende porphyry and andesite tuffs and flows. Alteration is dominated by an argillic assemblage of kaolinite, quartz, and pyrite, which is cross-cut by stockworks of quartz and chalcedony. Supergene mineralization occurs to depths ranging between 10 m to 80 m and consists of fracture-controlled and disseminated chalcocite. Rare native copper has also been discovered in this zone. Hypogene mineralization consists of chalcopyrite, pyrite, and rare bornite and covellite.

Petaquilla Copper drilled 195 holes in the Botija Abajo area between 2006 and 2008. Drill spacing is locally 25 m. Two zones of mineralization are present – Botija Abajo West and Botija Abajo East. Mineralization at Boitja Abajo West is dominantly copper rich with local Au-enriched zones. Its dimensions are 500 m x 350 m and it has been tested to depths of at least 140 m below the surface. The zone is open along strike to the northwest and southeast and at depth. Botija Abajo East is a southeast-trending zone that has a strike length of 650 m, width of 150 m, and extends to a depth of approximately 50 m. The Cu-Au mineralization is associated with a well-defined Au-Mo soil anomaly.

Relogging of the well-mineralized holes at Botija Abajo is recommended to help develop a geological model for this zone.

9.7 Medio

The Medio prospect lies between the Colina and Botija deposits. It was discovered by prospecting of Cu-Mo soil anomalies. Several drill holes in the area intersected Cu mineralization hosted in porphyry/granodiorite. MPSA is currently drilling this zone to assess its size and grade.




9.8 Cuatro Crestas

The Cuatro Crestas area is 2 km west of the Valle Grande deposit. The area is underlain by andesite and minor granodiorite intrusions. Stream and rock sampling in 1996 defined an area of anomalous Au and Cu values. Follow-up soil sampling outlined a Cu soil anomaly with dimensions of 600 m (east-west) x 1 km (north-south). Gold values are locally elevated in the soils in this zone. The soil anomaly is associated with a magnetic high. Five drill holes tested this zone in 1995, and Petaquilla Copper drilled another 23 holes in the area in 2007. Zones of elevated Cu and Au were intersected in drill holes over an area with approximate dimensions of 300 m x 200 m in the southern part of the soil anomaly.

9.9 Lata

The Lata area, 4 km northwest of the Colina deposit, is underlain by andesite tuffs and flows and argillic altered diorite and granodiorite intrusions. Gold soil anomalies occur on the edge of a magnetic high. Adrian Resources drilled eight holes in the area in 1995 and intersected anomalous Au values.

9.10 Nada

The Nada showing is 1.5 km west of the Molejón deposit. Drilling in the area tested Cu-Mo soil anomalies. Weak Cu-Au porphyry-style mineralization is hosted in andesite breccias.




10.0 EXPLORATION

Copper-gold-molybdenum porphyry style mineralization was discovered in central Panamá during a regional survey by the United Nations in 1968. Exploration by several companies has since outlined three large deposits and several smaller ones. An overview of the exploration history on the property has been presented previously in Table 6-1. Significant results of the historical exploration are presented below and additional information on the drill programs is presented in section 11.

10.1 Historical Surveys

10.1.1 Adrian (1992-1995)

Adrian Resources completed soil and auger geochemical surveys on most of the Concession. Line spacing was 200 m, with more detailed coverage (50 m or 100 m) around the known deposits (Figure 10-1). Between 1992 and 1995, approximately 8,000 soil samples were collected at depths of 5 cm to 20 cm. After drying on site, the samples were shipped to TSL in Saskatoon for analysis of copper, gold, molybdenum, and 30-element ICP. In addition, 3,600 auger samples taken at depths ranging from 50 cm to 90 cm were collected in areas of anomalous Cu and Mo. Contoured copper, gold, and molybdenum data for the concession are also shown in Figure 10-1. Details of these programs are presented in McArthur et al. (1995).

Several lithogeochemical sampling programs were also implemented, including reconnaissance-scale rock sampling (890 samples) and grid-based rock (172 samples) and soil (265 samples) sampling by Minnova, 1990-1992; detailed sampling in areas of suspected mineralization by Adrian; and reconnaissance sampling during the course of geological mapping and silt sampling in areas investigated for waste dump and tailings pond sites (Teck). The results indicate that near-surface copper mineralization is present in several areas, including the upper RÍo del Medio drainage and north of the Botija deposit.




Figure 10-1: Soil Geochemistry

Note: Soil survey area (blue outline), Mo >100 ppm (blue areas), Cu >400 ppm (green areas), Au >100 ppb (yellow areas).

10.1.2 Petaquilla Copper (2006–2008)

During the first half of 2008 Arce Geofisco of Lima, Peru, completed 105.2 km of IP surveying over the southeastern part of the Concession (Figure 10-2). The survey was done on north-south-oriented lines spaced at 200 m intervals using a pole-pole array with a spacing of 50 m and n=5.




Figure 10-2: Plan Map – IP Chargeability at -150 meter Level

A well-defined chargeability high is associated with the Botija deposit and the eastern edge of the Valle Grande deposit. A number of smaller IP chargeability anomalies occur along a southeastern trend between Botija and Botija Abajo.

10.2 MPSA (2007–2009)

Detailed geological mapping of the road cuts and drill pads in the Botija area was undertaken by MPSA geologists to resolve problems with structures defined by the 1996 mapping program. No evidence was found for the major structures suggested in the 1996 report (Speidel and Faure, 1996).




11.0 DRILLING

Since 1968 a number of drill programs have been conducted to test the extent of porphyry copper mineralization on the Concession. Details of these drill programs are summarized in Table 11-2. A drill hole location map is provided in Figure 11-1.

Table 11-1: Summary of Drilling by Operator and Area

Colina Botija Valle Grande Other Totals

Program No. of holes Metres

No. of holes Metres

No. of holes Metres

No. of holes Metres

No. of holes Metres

UNDP (s) 3 91.5 9 235.2 6 216.5 - - 18 543.2

UNDP (d) 4 670.8 4 728.4 2 412.4 - - 10 1,811.6

PMRD (s) 18 560.3 12 373.0 - - - - 30 933.3

PMRD (d) 30 7,236.2 20 5,199.3 1 207.1 - - 51 12,642.6

Adrian (1992-1995) 42 9,001.2 54 17,003.0 112 23,699.7 188 26,035.7 396 75,739.6

TeckCominco (1996)

47 12,718.1 32 8,243.8 12 1,879.1 33 3,995.6 124 26,836.6

TeckCominco (1997)

13 2,271.9 9 2,379.5 7 1,202.7 14 2,245.1 43 8,099.2

Petaquilla Copper (2006-2007)

4 268.2 38 1,472.1 43 2,153 230 26,575.1 315 30,468.4

MPSA (2007-2009) 65 18,611.7 105 33,765.1 47 11,220.8 71 9,863.5 288 73,480.7

Total 226 51,449.6 283 69,399.4 230 40,991.3 536 68,715.0 1,275 230,555.2

Table 11-2: Summary of Drilling by Area

Area No. of Holes Total Metres

Botija 283 69,399

Colina 226 51,450

Valle Grande 230 40,991

Botija Abajo 228 25,335

Brazo 22 6,480

Cuatro Cresta 28 4,060

Lata 8 1,081

Medio 21 3,926

Mestizo 5 776

Molejón 130 15,341

Nada 9 1,545

Tailings Area 44 4,255

Plant Sites 41 5,917

Total 1275 230,555




Figure 11-1: Drill Hole Location Map

Note: MPSA 2007-2009 (magenta); Petaquilla Copper 2006-2007 (green); Historical (pre-2007) (grey).

11.1 Historical Drilling

11.1.1 United Nations Development Program (1967-1969)

In 1968 and 1969, the UNDP drilled 27 short (average 30 m) holes with a Winkie drill (UNDP [s]). The exact locations of these holes can no longer be confirmed, and they were not used in this evaluation. Some of these holes remain in the database to aid exploration but are not used for resource estimation. The UNDP also drilled ten deeper holes with a BSS-1 drill in Botija, Colina, Valle Grande, and Medio (UNDP [d]). Core recovery averaged 70%. The locations of some of these holes have been confirmed, and in these cases the assay data remain in the database. No core exists from this phase of drilling, and geology from the drill holes was not used to develop the geological model.




11.1.2 Panamá Mineral Resources Development (PMRD) (Japanese Consortium)

Between 1970 and 1976, PMRD drilled 48 short holes (PMRD [s]) and 51 long holes (PMRD [d]) to test the extent of mineralization in the main deposit areas. The short holes were drilled with a Winkie drill, but no accurate location information or assay data exist, and these holes have been removed from the database. The long holes were drilled in the Botija and Colina deposits using a drill spacing of approximately 200 m. These holes have been re-surveyed. Core recovery averaged 95.5% (PMRD, 1977). Core recovery was poor in the surficial weathered intervals but excellent in unweathered rock. Geology is summarized on graphic logs at a scale of 1:1,000. No drill core, sample rejects, or pulps from this drilling remain.

11.1.3 Adrian Resources (1992 – 1995) (Inmet-Adrian-Georecursos)

As project operator Adrian Resources, a Vancouver-based junior company, completed 396 drill holes totalling 75,740 m between 1992 and 1995. The work was done by Falcon Drilling of Prince George, B.C., using three F-1000 hydraulic drills and one Longyear 38. Two Hughes 500D helicopters supplied by Coclesana SA and based at the Botija camp were used to move drills and crews.

Core recoveries were generally poor in overburden (20% to 80%) but very good, near 100%, in areas below the weathered horizon. Core diameter was thin-wall B (BTW) or NQ.

At Botija, 54 vertical drill holes were drilled, which reduced the hole spacing to 100 m. At Colina, 38 vertical holes tested the southwest gold zone and the main deposit on 200 m centres. At Valle Grande, 118 holes tested the mineralized zone at a drill spacing of 100 m to 200 m. Most of these holes were drilled with an azimuth of 220° or 40° and dip of -50° to cut across a hypothesized northwest-trending structural grain. Other drill holes completed by Adrian tested the Molejón Gold zone and other exploration targets on the mineral Concession.

Skeleton core from most holes is available and stored at the MPSA New Camp core storage facility. Most of the original core was left on site where termites and weather destroyed the boxes, rendering the core generally mixed and useless.

11.1.4 Teck (1996–1997)

In 1996 Teck commenced an infill drill program of 124 holes, for 26,837 m on the Botija, Colina, Valle Grande, and Molejón deposits. Specific objectives of this program were as follows:

define the global limits of mineralization at Botija, with emphasis on the east and southeast zones

locate the limits of and confirm potential at Colina by tightening drill hole spacing to approximately 100 m

evaluate starter pit potential and define the limits of mineralization in the northwest zone at Valle Grande.




The 1997 drill program was a continuation of the 1996 program, beginning in March of 1997 and lasting four months. The program was designed by Teck and Adrian and implemented and supervised by Simons. A total of 43 drill holes totalling 8,099 m were completed. The main objectives of the program were as follows:

collect metallurgical test samples from Botija and Colina test the continuity of higher-grade zones within Valle Grande for possible starter pit

potential complete exploration drilling in the Medio area east of Colina complete infill and step-out drilling on the Molejón Gold zone.

11.1.5 Petaquilla Copper (2006–2008)

From 2006 to 2008 Petaquilla Copper drilled 315 holes of HQ size totalling 30,468 m. Drilling was done using helicopter-supported Longyear 38 drills. Holes at Botija and Valle Grande assessed the potential for oxide copper mineralization. Drilling at Botija Abajo assessed the gold potential of this area. In addition, several exploration targets at Brazo, Cuatro Cresta, Lata, etc., were drilled.

11.2 MPSA (2007–2009)

During the period October 2007 to October 2009, MPSA drilled 288 HQ holes totalling 73,481 m on the Cobre Panamá property. Cabo Drilling Panamá Corp. provided the drilling services using a variety of drill machines. Most drill moves were done using helicopters provided by Heliflight Panamá S.A.

The objectives of this program were as follows:

provide sufficient drill density at Botija, Colina, and Valle Grande to calculate indicated mineral reserves

provide material for metallurgical and grinding tests from seven holes at Botija, five at Colina, four at Valle Grande, and one hole at Brazo

complete infill drilling at Botija to bring the reserves in the starter pit to the measured category. In addition, these holes were used for grinding tests to establish throughputs for this pit.

drill geotechnical holes at Botija and Colina

drill condemnation holes at possible plant site locations and in the tailings area.

Core recovery data from MPSA and Teck-Adrian drill holes located in the Botija and Colina deposit areas were reviewed. Values range from 1% to 100% with an overall average of 93%. Less than 5% of the sample intervals have recoveries below 50%, and 90% of the data have recoveries greater than 80%. There are no indications of any correlation between metal grade




and recovery. No data adjustments or exclusions related to core recoveries were carried out before the resource model was developed.

11.3 Surveying

11.3.1 Downhole Surveying

Historical Drilling

There is no record of downhole surveys for holes drilled before 1992.

During the 1992-1997 period, Adrian and Teck performed down-hole surveys on all holes using a Tropari device or acid tests. One test near the bottom of the holes was normally adequate for vertical holes, as readings rarely deviated by more than 1°. Inclined holes normally deviated with depth by a few degrees or more and required two to three tests per hole. Spurious results were discarded. Probable sources of error with the tropari tests include an abundance of magnetite in some areas and rock types and possible equipment malfunction due to corrosion in the tropical environment.

No downhole surveys were completed on the Petaquilla Copper drill holes. Most of the holes are vertical and shallow, and because the drilling was HQ, significant deviation is unlikely.

MPSA

Downhole surveys were completed on all geotechnical drill holes using a Reflex Maxibor II instrument. All resource drill holes greater than 300 m in depth were surveyed with the Maxibor instrument or the FLEXIT smart-tool single shot. Measurements with the FLEXIT were taken at 60 m intervals throughout the hole.

11.3.2 Collar Surveys

Geographic Projection

Before April 2009, all drill hole collars were surveyed in North American Datum 27 (NAD 27), Canal Zone. Subsequent to this date, all historical collar locations were converted to and any new holes were located using WGS 84, Zone 17N, which is the geographic projection for all engineering work on the project. WGS 84 coordinates are 19.7 m east and 207.1 m north of NAD 27, Canal Zone in the area of the Concession.

Pre-2006 Drill Programs

A number of drill holes completed by Adrian (i.e., drill collars to hole MO94-80) were located using a global positioning system (GPS) survey conducted by ECCCO Management of Squamish, B.C. ECCCO used a Trimble 4000 SE instrument and base station.

Teck subsequently surveyed most other holes in the Botija, Colina, and Valle Grande starter pit and Molejón deposit areas by conventional methods (total stations). Locations of many of the




holes that tested exploration targets are approximate, based on hip chain and compass traverses from known locations or on hand-held GPS readings.

Petaquilla Copper (2006–2008)

Petaquilla Copper used a hand-held GPS to survey drill hole locations. In late 2009 MPSA contracted GeoTi S.A., a Panamá City-based surveying outfit, to re-survey 46 of 177 holes drilled by PTC in the Botija Abajo and Brazo areas (19% of the entire PTC collar database). GeoTi carried out the work utilizing a Topcon HiPer Lite plus differential GPS system. After sorting out some geographic projection issues, MPSA considers this collar database to be validated.

MPSA (2007–2009)

During the 2007-2009 MPSA drill programs, drill collar locations were identified in the field using a GARMIN GPS-60CSx hand-held GPS unit. After each drill hole was completed, a cement monument with hole number and depth was constructed at the site. The collar locations were surveyed by GeoTi S.A. using a differential GPS system and base station (Topcon Hiper Lite plus). This system is accurate to 5 cm. All of the MPSA drill holes were surveyed.

In 2008 MPSA contracted GeoTi S.A. to survey 61 historical drill collars (11% of database). All collar coordinates were found to be within 5 m of original historical coordinates except for hole B96-33, which returned coordinates 10 m southwest of original site. Almost all coordinates in the 2008 survey were shifted southwest from original survey. The reason for the shift in the GeoTi data has not been resolved.

MPSA checked the location of another 29 holes using a hand-held Garmin GPS 60Csx during the 2007-2009 field programs. All locations were validated except Botija hole LBB-038, which was located in the field 33 m west and 35 m south of the original historical surveyed collar coordinate. This hole was corrected in the database.




12.0 SAMPLING METHOD AND APPROACH


12.1.1 UNDP

No record of the sampling procedures used by UNDP remains. The record indicates that UNDP took samples at 3 m intervals.

12.1.2 PMRD

PMRD samples were half-split core using a mechanical splitter. This was an industry standard procedure at the time and is still used occasionally. PMRD samples were uniform 2 m in length.

12.1.3 Adrian, Teck

Adrian and Teck sample intervals were 1.5 m in length from the collar of the holes. Samples were marked and split using a mechanical splitter. Teck metallurgical samples average 4.5 m and the length of the samples was modified by lithology. According to the record, metallurgical samples were not split, but entirely consumed for testing.

12.1.4 Petaquilla Copper

The following procedures were in place during the Petaquilla Copper drill programs:

Core is delivered to the logging facility at Colina camp by helicopter. Core boxes are closed at the drill and then transported in specially constructed cages slung below the helicopter to eliminate spillage from the boxes during transport. Cages are deposited at the helicopter pad and transported by truck and workers to the core logging facility. The core is photographed and then placed on tables in the logging facility, where it is geologically logged and the sample intervals are marked. Sample intervals are uniformly 1.5 meters in length and disregard lithological boundaries. A sample is selected approximately every metre for specific gravity determination, which is performed by weighing the sample in air and then in water. The samples are dried prior to immersion, but not coated with a sealing agent.

After the specific gravity determination is completed, the sample is returned to the core box. When logging is completed, the core is moved to the core splitting area, where it is split using rock saws with specially constructed jigs that hold the core securely and insure that the core is split in the centre of the core. One half of the core is then replaced in the core box for archive and the other half is placed in a large plastic bag marked with the sample number. The bags containing the samples are then transferred to the sample preparation facility. The archive samples are placed in racks in the storage facility.

12.2 MPSA

The drill contractor places the HQ drill core into wooden core boxes (1 m long x 3 rows) at the drill rig. Lids are placed on the boxes and secured by wire before transport by helicopter to the




drill contractor’s staging area. Core is transported by truck to the Colina core facility by workers supervised by MPSA.

The core boxes are washed and laid out in order, and meterage blocks are checked. High-resolution digital photos are taken starting at the top of the hole. Each photo shows two boxes and a sign with the hole number, interval, and a colour bar. Images are downloaded onto the site computer and relabelled with the hole and core box numbers. The JPEG files from each drill hole are combined into one PDF file and included in the digital filing cabinet.

The drill core is placed on the core logging benches where it is examined by MPSA geologists who prepare geological and geotechnical paper logs for each hole. The paper logs for each hole are scanned as PDF documents which are included in the digital filing cabinet. The original logs are stored at site. These logs are entered into Excel spreadsheets by a data entry assistant. These Excel files are subsequently imported into the ACCESS database.

MPSA geologists and/or geological assistants select the assay samples. Sample intervals are 1.5 m except where they are modified due to major lithological changes. Bar-coded sample tags are then placed in the boxes at the start of each sample interval. An aluminum metal tag with the sample number is attached to each box by geological assistants. The sample numbers, hole number, and intervals are entered into an Excel spreadsheet.

In the Colina core saw room, MPSA personnel cut the drill core in half using production rock saws. Saprolite is sampled with a spoon or sheet-metal tool similar to a trowel. One half of the core sample is returned to the boxes for storage at the new camp facility, and the other half is placed in a plastic bag with the bar-coded sample ticket. These bags are taken to the sample preparation laboratory.




13.0 SAMPLE PREPARATION, ANALYSES AND SECURITY


13.1.1 UNDP

No record of the sampling preparations, analyses and security used by UNDP remains. No core remains from this phase of drilling and geology from the drill holes was not used to develop the geologic model.

13.1.2 PMRD

The core from the PMRD drill program was split with a mechanical splitter. Samples were taken at 2 m intervals down the hole. The rock was pulverized and then sent to the assay lab of Direction General de Recursos Minerales of Panamá where they were analyzed for Cu and Mo using an atomic absorption instrument. Thirty duplicate samples were sent to the Central Research Lab of Mitsui Smelting and Mining Limited for check assaying. The Direction General de Recursos Minerales of Panamá assays were slightly lower, but within acceptable limits. The Mitsui results averged 0.597% Cu and the Panamá results averaged 0.578% Cu. Results of this work were deemed satisfactory.

Adrian and Teck twinned two holes for grade comparison (Simons, 1996). Areas with adjacent holes were also identified in the western part of Colina and Botija; some differences were noted, but the PMRD data were generally validated by the nearby holes.

13.1.3 Adrian, Teck

Sample Preparation

The following procedures were in place during the 1992-1995 Adrian drill programs:

After delivery to the core logging and sample preparation facility at Colina, the core was marked into 1.5 m lengths. Geological logging was done by Canadian geologists hired by Adrian and geotechnical logging by Panamanian geologists hired by Geotec. A modified form of GEOLOG was used to record geological data, including lithology, alteration, mineralization, and structure. Geotechnical information included recovery and RQD values for each 1.5 m interval.

Adrian and Teck sample intervals were 1.5 m in length from the collar of the holes. Samples were marked and split with a mechanical splitter. Half of the split core was archived, but this material has been lost. The remaining half was crushed and split with a Jones splitter. A one-eighth split weighing approximately 250 grams was taken for each 1.5 m interval for analysis. The equipment was cleaned with an air hose between samples. Core rejects were stored in plastic bags, which were stored on site.

Teck’s metallurgical samples average 4.5 m in length, although locally the length was modified if there were lithological changes. According to the record, metallurgical samples were not split but entirely consumed for testing.




Assaying

Samples from the Adrian and Teck drilling programs were analyzed at TSL Laboratories in Saskatoon, Canada. The general procedures are as follows:

Copper analysis was performed by acid digestion of a 0.5 gram sample and analyzed by AA (atomic absorption).

Molybdenum analysis was performed by acid digestion of a 1 gram sample and analyzed by AA.

A 30 gram sample was used for gold analysis by fire assay with an AA finish.

Silver analysis was performed by acid digestion of a 2 gram sample and analyzed by AA.


Sample Preparation

At the sample preparation facility, the samples are transferred from the plastic bags to metal trays and placed in an electric drying oven at 115°C for three to six hours, depending on the nature of the material. Temperature in the oven is electronically monitored and maintained. After the samples are dry, they are immediately sent to the crusher (a small BICO jaw crusher) where the entire sample is crushed to approximately 5 mm to 6 mm. The crusher is cleaned after each sample by crushing a small amount of blank basalt and by blowing remaining material out of the crusher with compressed air. The process takes three to six minutes depending on the hardness of the sample.

In the event of failure of the power supply or problems with the electric oven, the facility has two wood-fired drying rooms, where wood fires in stoves are used to heat a room to approximately 60°C. Samples require more drying time than in the electric oven, but the capacity is significantly larger than the electric oven, thus insuring that a continuous stream of samples is available to the crusher.

A 250 gram aliquot of sample is then split using a riffle splitter with approximately 12 mm (½ inch) openings. The reject material is replaced in plastic bags, which are then clearly marked and stored in a covered storage area on site. The 250 gram aliquot is placed in a small plastic bag with a sample tag and heat-sealed. The plastic bags are very heavy and durable, and tamper evident when sealed properly. Approximately 60 samples are then placed in large woven plastic (rice) bags, which are tied off and clearly marked. Samples from different parts of the deposit are sent to different laboratories, and the destination is clearly marked at the sample preparation facility.

The sample preparation facility is locked when no one is working there, and when the doors are not locked, someone is always present in the facility. This policy is in place to monitor the integrity of the samples.




Samples were trucked on Mondays and Wednesdays to the Petaquilla Copper office at La Pintada, where they were placed in sturdy boxes and labelled for shipment. The boxes were immediately placed in a truck for transport to Panamá City, where they were entrusted to FedEx for shipment to SGS Laboratories in Lima, Perú, or ALS Chemex in Vancouver, Canada. The samples arrive in Panamá City the same day they are shipped from the sample preparation facility.

Assaying

Samples sent to SGS Laboratories in Lima, Perú, were analyzed for sequential copper and for Au, Ag, Cu total, and Mo using the standard assay techniques. Samples sent to ALS Chemex in Vancouver were analyzed for Cu, Au, Ag, Mo, and multi-element ICP using standard assay techniques.

13.2 MPSA

Sample Preparation

MPSA personnel placed samples in aluminum trays which were transferred to ovens where they are dried for 12 hours at 90°C. The entire sample is then crushed to -10 mesh (2 mm) using a Rocklabs Boyd crusher. Regular (at least twice a day) sieve tests are carried out to ensure that material is being crushed to the appropriate size. If the quantity passing falls below 80%, crusher jaws are adjusted accordingly. A written record of this test is available for review. The crusher is cleaned with high-pressure air after every sample. After every 10 samples a coarse blank sample is passed through the crusher.

Each crushed sample weighs approximately 8 kg. This material is split using a Jones riffle splitter. A 500 gram aliquot of each sample is taken for assay and placed and heat-sealed in a small plastic bag marked with a bar-coded sample tag. The remaining material is returned to the original sample bag and stored on site. The assay samples are shipped by air courier to ALS Chemex in Lima, Perú, for analysis.

Assaying

Samples collected from 5 October 2007 to 1 October 2009 were assayed by:

ALS Chemex Lima Calle 1 LT-1A Mz-D Esq. Con Calle A Urb. Industrial Bodanegra Callao 1 Lima, Peru

Check and other secondary assay work was conducted by:

Acme Santiago S. Av. Claudio Arrau 7152, Pudahuel Santiago, Chile

Both labs have ISO 9001:2000 certification.




Sample Reduction and Analytical Procedures

After receiving the samples, ALS Chemex scans the bar code on the assay sample bag to enter the sample into its system. Half of the sample (250 grams) is pulverized with a LM5 pulverizer with low-chrome steel parts. This results in 85% of the material being less than 75 µm in size. Copper assays are by AAS after a four-acid digestion (HF-HNO3-HClO4-HCL). Gold analyses are by fire assay and AAS analysis on a 30 gram sample. Silver and molybdenum are included as part of multi-element ICP-AES scan. Total sulphur is analyzed using a LECO induction furnace.

Near-surface samples are analyzed for sequential copper to estimate the amount of leachable copper. Sulphuric, cyanide, and citric-acid soluble copper and residual copper values are reported for each sample.

All pulps are currently being stored at ALS Chemex in Lima, Perú.

Secondary Laboratory Analyses

Approximately 1 in every 22 samples is chosen randomly for check analysis at Acme Analytical Labs in Santiago, Chile. A sample list is forwarded to ALS Chemex in Lima, which selects the requested sample pulps and sends them to Acme in Santiago. These samples are analyzed for copper and gold using the same analytical methods as ALS Chemex. Results are cross-referenced with original ALS-Lima results and sent to BD Resource Consulting in Denver, Colorado, for QA/QC analysis.

Sample Security – Chain of Custody

The assay samples are kept in a pad-locked facility until they are ready for shipment. Samples for each hole are shipped out only when all samples from the hole are complete. Samples are collected into larger bags in batches of 90 individual samples. Sample intervals being dispatched for sequential copper analysis are sent out in separate shipments of 20 to 25 samples. Every three or four days the large bags are driven by MPSA personnel to a secured MPSA warehouse in Penonomé, where they are kept under lock and key until picked up by DHL cargo shipping. The samples are usually in the warehouse for less than two days. DHL airfreights the samples to ALS-Chemex Labs in Lima, Perú.

During a site visit (April 15-16th, 2009) the principal author (William Rose) and the QA/QC consultant (Bruce Davis) noted that all procedures described above were being carefully observed and meet or exceed industry standards for collection, handling and transport of drill core samples.




14.0 DATA VERIFICATION


No QA/QC programs were in place during the early drilling on the project (UNDP, PMRD) as this was not standard practice at the time.

14.1.1 Adrian, Teck

The Adrian QA/QC program in 1992-1995 consisted of a few check assay samples sent to XRAL and Chemex. These data were reviewed by Francois-Bongarcon (1995), who concluded “… which tends to support an absence of bias between laboratories for copper and a reasonably good reproducibility of pulp assays...” and that there is a strong bias between TSL and Chemex Au analyses. AMEC (2007) concurred with that assessment of the data but noted that few check assay data and supporting standards, and no duplicate data, were available.

During the 1996-1997 Teck programs, a combination of check assays was sent to Chemex Labs Ltd. in Vancouver along with standard samples to monitor the quality of analyses. Check assay samples were assayed for Cu, Au, Ag, and Mo. The standards were inserted into the sample stream at a rate of 1 in 15 samples. A discussion of the results of the check and standard assay results is given in the feasibility reports by Simons (1996, 1998).


QA/QC Sample Insertion

Duplicate and standard samples are inserted in the sample stream to assess the variability and accuracy of the assay results. A duplicate sample pulp is inserted approximately every 15th sample, and a certified standard sample obtained from Ore Research and Exploration Australia is inserted approximately every 20th sample. During the 2006-2008 drill program approximately 4.5% of the samples were duplicates and 3.2% of the samples were standards.

Control Sample Performance

The Petaquilla Copper quality control sample data were reviewed by Bruce Davis, F.AusIMM of BD Resource Consulting in Denver, Colorado, and found to be acceptable. Control charts indicated performance was mediocre. Coarse reject data indicate that sample preparation contributed to the precision problems; however, the copper numbers do not indicate the presence of any bias in the assays and the program may be used in conjunction with other drilling information from other campaigns.

Control samples were inserted with the samples collected during the drill campaign conducted by Petaquilla Copper S.A.. The copper assay results of standard reference material indicate the copper assays from the program are valid as shown in Figure 14-1.

.




Figure 14-1: Results for ORE 52Pb – Reference Material Certified for Cu

Gold results were more erratic than copper over the range of values likely to be encountered in the deposit areas. An example with a slightly higher grade standard is shown in Figure 14-2. Generally, the performance on the higher grade standards was acceptable; accuracy was good while precision is as indicated in the figure.




Figure 14-2: Results for ORE 53 Pb - Reference Material Certified for Au

The blanks showed no indication of contamination in the samples (Figure 14-3, 14-4).

Figure 14-3: Blank Material - Cu%




Figure 14-4: Blank Material - Au ppm

In addition, 10 holes from Botija Abajo (7 holes) and Brazo (2 holes) were reassayed using the archived core. One hole, BRDH-005, was re-drilled as BR09-019M. Analytical results of this work are comparable to the original assays done by Petaquilla Copper. QA/QC performance was good for this work and is described as part of the MPSA work program.

14.2 MPSA

QA/QC Sample Insertion

QA/QC sampling was initiated at the start of the FEED drill program in October 2007. A series of blank samples, crusher duplicates, prepared standards, and core duplicates were inserted at random into the sample stream being sent to the analytical lab. In addition, a proportion of the samples were re-assayed for Cu and Au at a second lab. During the initial phase of the program (October 2007 to June 2008) the percentages of QA/QC samples are as follows: blanks (4.7%), crusher duplicates (4.9%), prepared standards (5.0%), core duplicates (4.1%), and umpire/check analyses (5.0%). After a review of these data in July 2008, the proportion of QA/QC samples was reduced to the following: blanks (3.1%), crusher duplicates (3.2%), prepared standards (4.0%), and umpire/check analyses (4.5%). Core duplicates (¼ cores) were discontinued after the initial phase because enough of this type of QA/QC sample had been collected to alleviate any concerns. There was sometimes high variability in core duplicates, especially with gold results, often due to geological variability in the sample, veins or a nugget effect.




Blanks

Blank material was inserted at a rate of roughly 1 per 30 samples. The blank material is derived from an unmineralized granodiorite outcrop adjacent to the bridge over RÍo San Juan. The material is visually identical to the Petaquilla batholith. Blanks are inserted into the sample stream at the crusher and therefore have the dual purpose of cleaning the jaws and serving as a control on potential contamination.

Standards

Commercially available certified porphyry standards from CDN Labs in Vancouver, Canada, and Ore Research and Exploration (OREAS) of Melbourne, Australia, have been used. Standard samples are inserted at random into the sample stream at the Colina sample preparation facility.

Review of QA/QC Data

All QA/QC samples results after July 2008 were forwarded for independent review by geostatistician Bruce Davis (BD Resource Consulting Inc.) of Denver, Colorado.

The analysis includes all results from 2007 through the holes drilled for grinding studies in 2009. Only example graphs are given with the text. A complete set of graphs for SRM, coarse reject duplicates, umpire duplicates, and blanks is provided in a separate document. Note that the charts are made with uncorrected data. Comments are given in the text explaining out of control results. In some cases explanatory annotation is also provided on the chart. The results were divided into two groups, 2007-08 and 2009. The division is merely a matter of convenience and it does not imply the program was significantly different between years.

Standard Reference Material (SRM) Performance

Results for the copper standards fall within the control limits above the prescribed rate (Figure 14-5). In almost all cases, values falling outside of the prescribed control limits are the result of mix ups in control sample insertion into the sample stream. There appear to be very few if any errors generated by the assay process. Most of the mis-assignment errors have been reviewed and appropriate changes made to the data when it was clear the sample was not the standard expected; however, a few sample results remain (well less than one percent) where no explanation for the discrepancy is available.




Figure 14-5: Results for CM-1 - Reference material certified for Cu

3002001000

Sequence Number

1.30

1.20

1.10

1.00

0.90

0.80

0.70

0.60

0.50

0.40

Cop

per

(%)

Cobre de Panama 2007 and 2008

SRM CM-1 Copper

AVG = 0.8672

LCL = 0.768

3002001000

Sequence Number

1.30

1.20

1.10

1.00

0.90

0.80

0.70

0.60

0.50

0.40

Cop

per

(%)

Cobre de Panama 2007 and 2008

SRM CM-1 Copper

AVG = 0.8672

LCL = 0.768

UCL = 0.938Accepted value = 0.853

13 of 257 are out of control

3002001000

Sequence Number

1.2

0.9

0.6

0.3

0.0

Cop

per

(%)

Cobre Panama 2009

SRM CM-1 Copper

AVG = 0.8258

LCL = 0.768

UCL = 0.938Accepted value = 0.853

The results for gold also fall within the control limits at or above the prescribed rate of ninety percent (Figure 14-6).




Figure 14-6: Results for CM-1 - Reference Material certified for Au

3002001000

Sequence Number

3

2

1

0

Gol

d (g

/t)

Cobre Panama 2007 and 2008

SRM CM-1 Gold

AVG = 1.8457

LCL = 1.67

UCL = 2.04

Accepted value = 1.85

3002001000

Sequence Number

2.50

2.30

2.10

1.90

1.70

1.50

Gol

d (g

/t)

Cobre Panama 2009

SRM CM-1 Gold

AVG = 1.8114

LCL = 1.665

UCL = 2.03

Accepted value = 1.8520 of 262 out of control




Sample Blank Performance

Blank results for Cu and Au are shown in Figure 14-7 and 14-8 respectively. Only what appear to be sample “swaps” in the copper blank material exceed the control limit. The out of control samples have been reviewed in the context of the affected batches. No evidence of contamination was ever found.

Figure 14-7: Blank Material - Cu%

8006004002000

Sequence Number

0.40

0.30

0.20

0.10

0.00

Cop

per

(%)


Copper Blanks

AVG = 0.0127UCL = 0.03

10008006004002000

Sequence Number

0.3

0.2

0.1

0.0

Co

pp

er

(%)

Cobre Panama 2009

Blank Copper

AVG = 0.0068

UCL = 0.03

5 of 942 failures

Gold has a few more “out of control” results but the number of out of control results still does not exceed the prescribed rate for outliers. There is no evidence of contamination in the samples.




Figure 14-8: Blank Material - Au g/t

8006004002000

Sequence Number

0.12

0.10

0.08

0.06

0.04

0.02

0.00

Gol

d (g

/t)


Gold Blanks

AVG = 0.0064

UCL = 0.02

10008006004002000

Sequence Number

0.08

0.07

0.06

0.05

0.04

0.03

0.02

0.01

0.00

Gol

d (g

/t)

Cobre Panama 2009

Blank Gold

AVG = 0.0060

UCL = 0.02

7 of 942 exceed control limit

Coarse Duplicate Sample Performance

Coarse duplicate samples check the adequacy of the sample preparation protocol. The charts below (Figure 14-9, 14-10) are for example batches which show the pairs fall within control limits at above the prescribed rate. This indicates the sample preparation process is under control and




producing samples that are sufficiently homogeneous to form the basis for resource estimation. A relatively large percentage of the gold values fall outside of the control limits at the lower end of the grade range. This appears to be due to assay detection limit effects rather than a problem with the preparation process.

Figure 14-9: Coarse Reject Duplicate Results - Cu%

3210

Original Copper (%)

0.9

0.6

0.3

-0.0

-0.3

-0.6

-0.9

Rel

ativ

e D

iffer

ence

Cobre Panama 2007-2008

Coarse Reject Duplicate Results

+30%

-30%

More than 95% of pairs are within 30%

3210

Original Copper (%)

0.9

0.6

0.3

-0.0

-0.3

-0.6

-0.9

Rel

ativ

e D

iffer

ence

Cobre Panama 2009


+30%

-30%

Over 95% within 30%




Figure 14-10: Coarse Reject Duplicate Results - Au g/t

0.60.50.30.20.0

Original Gold (g/t)

0.9

0.6

0.3

-0.0

-0.3

-0.6

-0.9

Rel

ativ

e D

iffer

ence

Cobre Panama 2007-2008


+30%

-30%

Over 90% within 30% for Au > .07

0.80.60.40.20.0

Original Gold (g/t)

0.9

0.6

0.3

-0.0

-0.3

-0.6

-0.9

Rel

ativ

e D

iffer

ence

Cobre Panama 2009


+30%

-30%




Pulp Duplicate Sample Performance

Check assays were performed by Acme Analytical Labs in Santiago, Chile. Results for copper indicate the checks validate the original assay results as shown in Figure 14-11.

Figure 14-11: Interlaboratory Checks - Cu%

Pulp Duplicate Copper Assays

0.00

0.20

0.40

0.60

0.80

1.00

1.20

1.40

1.60

1.80

2.00

0.00 0.20 0.40 0.60 0.80 1.00 1.20 1.40 1.60 1.80 2.00

Original ALS Cu (%)

Du

pli

cate

Acm

e C

u (

%)

1.51.20.90.60.30.0

Copper (%)

0.5

0.4

0.3

0.2

0.1

0.0

-0.1

-0.2

-0.3

-0.4

-0.5

Re

lativ

e D

iffe

ren

ce

Petaquilla Copper Umpire Duplicates

ALS Original and Acme Duplicate

+10%

-10%




3210

Copper (%)

0.9

0.6

0.3

-0.0

-0.3

-0.6

-0.9

Rel

ativ

e D

iffer

ence

Cobre Panama Umpire Set #4

Copper Pulp Duplicate Differences

+10%

-10%

The gold results show a similar pattern of dispersion as the coarse reject duplicates (Figure 14-12). It appears that the results are unbiased (an average relative difference near zero), but results are erratic for values less than 0.15 g/t Au. The differences are most likely the result of detection limit effects in the assay process and are not an indication of any problem.




Figure 14-12: Interlaboratory Checks - Au g/t

Gold Duplicate Relative Differences

-0.5

-0.4

-0.3

-0.2

-0.1

0

0.1

0.2

0.3

0.4

0.5

0 0.5 1 1.5 2 2.5

Original ALS Au (ppm)

Rel

ativ

e D

iffe

ren

ce w

ith

Du

pli

cate

1.00.80.60.40.20.0

Gold (g/t)

1

1

0

-0

-0

-1

-1

Rel

ativ

e D

iffer

ence

Cobre Panama Umpire Set #4

Gold Pulp Duplicate Differences

+10%

-10%




Conclusions

Results from all types of QC samples indicate sample “mix ups” occur occasionally; however, the standard reference (SR) material indicates the assay process is producing valid results. Blank assays show that there has been no contamination in the preparation or assay process. The coarse reject work demonstrates the sample preparation process is well controlled. Assay from pulp duplicates submitted to Acme for assay confirm the values in the original assays.

The Cobre Panamá sampling and assaying program produces sample information that meets industry standards for copper and gold accuracy and reliability. The assay results are sufficiently accurate and precise for use in resource estimation and the release of drill hole results on a hole by hole basis.

There are no quality control checks applied to molybdenum or silver assays. The lack of control is considered to be of no consequence since these metals contribute comparatively little to project revenue.

14.3 Database Validation

A series of 44 holes were randomly selected from the MPSA database in order to conduct manual validation of the underlying database. Data was compared back to a master database provided by MPSA which includes scans of signed assay certificates and scans of original drill hole logs. A total of 43 errors were found for an error frequency of 0.55%. Eleven of these errors occurred in one drill hole in which the assay results were offset by three sample intervals. The remaining errors were primarily related to incorrect manual keying of data generated by Teck and Adrian in the mid-1990’s. None of the errors identified during this review are considered significant and would have negligible effects on the resource estimate. An overall error rate of less than 1% is considered to have virtually no affect on the overall mineral resource estimate.




15.0 ADJACENT PROPERTIES

There are no adjacent copper-producing properties. The closest significant copper property is the Cerro Colorado porphyry copper deposit located 120 km to the west.

15.1 Molejón

The Molejón Gold deposit is located approximately 4 km south of the Botija deposit and is owned by Petaquilla Minerals Ltd. of Vancouver, Canada. Preproduction began in July 2009.

The Molejón area is underlain by Tertiary volcanics and subvolcanic andesites, which have been intruded by feldspar-quartz porphyry dikes. Significant gold grades are associated with three northeast-trending structurally controlled zones of quartz-carbonate breccias and adjacent altered zones that dip at a shallow angle 25° to the northwest. The main near-surface mineralized zone is approximately 10 m thick. The deposit is interpreted to be a low-sulphidation, quartz-adularia epithermal gold deposit. Main zone veins locally exhibit typical banded cockade textures, and gold occurs as electrum. Economic mineralization consists of the oxidized portion of the breccia, feldspar quartz porphyry, and feldspar andesite flows.

On 18 October 2007 Petaquilla Minerals Ltd. announced a “NI 43-101 compliant gold resource estimate” completed by AAT Mining Services. Measured, Indicated, and Inferred resources were calculated at a cutoff grade of 0.5 g/t Au. Results are summarized in Table 15-1. These resources were taken from the Petaquilla Minerals website (www.petaquilla.com) and are provided for information only.

Table 15-1: 2007 Resource Estimate – Molejón Gold Deposit (Petaquilla Minerals)

Class Tonnes Au Grade (g/t) Ounces

Measured 9,445,109 1.75 532,801

Indicated 6,281,558 1.11 223,785

Inferred 12,209,004 0.872 342,111




16.0 MINERAL PROCESSING AND METALLURGICAL TESTING

16.1 Metallurgical Testing

16.1.1 Introduction


In 1997, an extensive program of metallurgical testing was designed to confirm earlier work on the metallurgical response of the Botija and Colina deposits. Most of the work was done at Lakefield Research Ltd., Lakefield, Ontario. Grinding, flotation, dewatering, and mineralogical work were performed as part of this program. In addition to the Lakefield work, locked-cycle flotation testwork and modal analysis were performed at G&T Metallurgical Services Ltd., Kamloops, B.C. (G&T) to assist in defining grind requirements for both rougher and cleaner flotation. Copper-molybdenum separation using differential flotation was conducted by International Metallurgical and Environmental, Kelowna, B.C. (IME). The metallurgical work done for the present study has built upon the 1997/1998 study with some knowledge of, but no reliance on, work performed before that time.

The testwork before 2007 was based on large composite samples, and the results, particularly for flotation testing, could not be used for interpreting the variability of response for material within the deposits. Consequently, a large sampling program was undertaken in 2008/2009 to bolster the knowledge from previous work and provide the missing insight into the variability of response. A total of 16 special holes for metallurgical grinding and flotation tests were drilled in the Botija, Valle Grande, and Colina orebodies. Sample preparation, flotation testing, and testing of flotation products were done primarily at G&T. Grinding work was conducted at SGS Mineral Services, Lakefield, Ontario, and at Philips Enterprises LLC, Golden, Colorado.

The program resulted in:

additional geological data a comprehensive suite of grindability parameters resulting in new throughput

estimates additional flotation response data for estimating concentrates production and

operating costs sample materials for marketing purposes additional design data for solid-liquid separation, regrinding, and pipeline design.

The resulting life-of mine production data are shown in Table 16-1.







t/a 71,968,053 73,281,400 71,421,733

t/d 197,173 200,771 195,676

Head Grade % Cu 0.46 0.34 0.41

% Mo 0.008 0.006 0.008

g/t Au 0.08 0.05 0.07

oz/t Ag 1.46 1.39 1.43

Recovery % Cu 88.4 79.5 85.9

% Mo 61.9 53.1 59.0

% Au 57.2 44.6 54.3

% Ag 47.3 42.8 45.8


% Cu 28 28 28


t/a 7,090 4,741 6,188

% Mo 52 52 52

16.1.2 Samples

For the 2008/2009 sampling campaign, a series of 11 drill holes, subsequently expanded to 12 holes, were sited in the Botija and Colina mineralized zones. The positions of the holes were selected to cover an even spread along the major axes (the centrelines) of the two zones; the holes were targeted to penetrate the major combinations of lithology and alteration identified in the 1998 geological model.

16.1.3 Grindability Testwork

A large amount of grindability data was collected in 2009 to supplement earlier work. The 2009 data confirmed preliminary assumptions of throughput rates and indicated that there were benefits to adding a pebble-crushing circuit to each grinding line.

16.1.4 Flotation Variability Testwork

Variability flotation test results indicated that the metallurgical response is variable but consistent across the Botija/Colina deposits. The main source of variability in copper recovery is the copper content in the feed. Lower average copper recovery for the Valle Grande master composites is related to lower levels of copper sulphide liberation. In general, copper losses in the rougher circuit were higher for the low- and medium-grade Valle Grande composites than for the Botija/Colina samples.

16.1.5 2009 Pilot Plant

In September 2009, three composite samples of ore were treated in a pilot plant at the G&T laboratories in Kamloops, B.C. The objectives of the pilot plant were to generate large




quantities of samples for further testing, primarily for molybdenum separation tests. In turn, the molybdenum separation tests generated samples for additional tests.

The pilot plant produced about 70 kg of final bulk concentrate, which was used to carry out bench-scale copper-molybdenum separation work and solid-liquid separation tests. The bulk concentrate produced in the pilot plant contained 27.6% copper and 0.51% molybdenum.

Copper-molybdenum separation tests were carried out on a blended bulk concentrate produced in the pilot plant. It was possible to produce marketable-grade molybdenum concentrates at about 90% molybdenum recovery.

Deleterious minor element concentrates in both the bulk and molybdenum concentrates were generally below penalty limits. Lead levels in the molybdenum concentrate at 0.04% may be on the threshold of penalty limits at some roasters.

16.2 Mineral Processing

Ore from the Botija, Colina, and Valle Grande pits will be treated in a concentrator to produce a copper concentrate and a molybdenum concentrate for sale on the world market. Initially, the concentrator will treat nominally 150,000 t/d of ore supplied from the Botija pit; later, ore will be received from the Colina and Valle Grande pits. From Year 10, the concentrator ore throughput will be increased by 50%, to a nominal 225,000 t/d, to maintain production of concentrate despite a falling head grade. Crushing, grinding, bulk rougher flotation, water, and air systems will increase in capacity by 50% to accomplish the increase in ore treatment rate; all other systems will remain at the same size.

The process plant is designed to process ore at a head grade of 0.7% Cu and 0.013% Mo. These levels are higher than the highest sustained head grades of 0.58% Cu and 0.011% Mo expected to be mined in Year 5, but the design provides the flexibility to accommodate a wide range of head grades over the project life. The plant design also allows for 15% day-to-day fluctuations in throughput. The process includes the following facilities:

crushing and grinding to liberate minerals from the ore froth flotation to separate most of the copper and molybdenum minerals from

minerals of no commercial worth differential flotation to separate the copper and molybdenum minerals from each

other facilities to store tailings and provide reclaim water for the process facilities to remove water from the products and to ship concentrates to market.

A simplified flow diagram is provided in Figure 16-1.




Figure 16-1: Process Flow Diagram

Primary Crusher

Open Pit Mine

SAG Mills (2)

Screens

Coarse Ore

Stockpile

Rougher Flotation (4 banks of 7)

3rd Cleaner Flotation (4)

2nd Cleaner Flotation (6)

1st Cleaner Flotation (2 banks of 8)

1st Mo Clnr

Flotation

Cu Conc.Thickener

Cu ConcStorage Tank

(2)

MainlinePumps (2x5)

Copper Filters (4)

Cu Conc. Loadout

Pipelines (32 km)

Tailings to Impoundment

Regrind Vertimills

(4)

Cu Conc.Storage Tank (2)

Ball Mills (4)

Cyclone Cluster (4)

CycloneCluster (2) Bulk Conc.

Thickener

Bulk Concentrate

Storage Tanks (2)

Mo Conditioner

o/f

Potentially Acid-Generating Tailings to Impoundment

To Process Water

Shiploader

Mo Rougher Flotation Mo Scav.

Flotation

2nd Mo Clnr Flotation

3rd Mo Clnr Flotation

4th Mo Clnr Flotation


Mo Conc.PackingMo Dryer

BinsMoFilter

Regrind Vertimill

Pebble Crushers (2)

Mo Concentrate Thickener

Thickener/Clarifier

Overflow

Port

MineSite

Run-of-mine (ROM) ore will be delivered by haul truck to the dump pockets of two primary gyratory crushers installed in a single in-ground concrete structure close to the rim of the Botija pit. A 400,000 tonne ROM stockpile will be located close to the crushers to provide a 2½-day supply of ore for times when weather conditions preclude hauling ore out of the pit. The ROM stockpile will be operated on a first in, first out basis to prevent the accumulation of aged ore.

Separate feeders and take-away conveyors will move the ore from each crusher to a series of conveyors which will discharge the ore onto a conical coarse ore stockpile at the concentrator. Provision will be made at the transfer point between the two overland conveyors to accept mill feed from future crushed ore sources. The coarse ore stockpile will hold a 2½-day supply for




the mill, 15 hours of which will be available to the reclaim feeders without the assistance of a bulldozer.

Two trains of feeders and conveyors will draw ore from below the coarse ore stockpile and feed two parallel wet-grinding lines, each consisting of a semi-autogenous grinding (SAG) mill and two ball mills, all equipped with gearless drives. The SAG mill circuits will be closed by trommel screens followed by washing screens; conveyors will deliver screen oversize to pebble crushers. The pebble crushing circuits will include pebble bins, cone crushers, and a bypass arrangement. Crushed pebbles will return to the SAG mills via the feed conveyors. From Year 10 of operation, another coarse-ore stockpile and grinding line will be added to increase in ore treatment rate.

Discharge from each SAG mill will be evenly split between two ball-mill circuits. The four ball-mill circuits will be closed by hydrocyclones. Ground slurry will be directed to a flotation circuit where a bulk sulphide concentrate, containing copper, molybdenum, and gold values, will be collected and concentrated in a rougher followed by three stages of cleaner flotation. The roughers and first cleaners will be tank cells, while the second and third cleaners will be column cells. Before cleaning, rougher concentrate will be reground in vertical stirred mills. From Year 10, a 50% increase in rougher capacity will be required to accommodate the increase in throughout, but the amount of copper will be the same; therefore, no change to the existing downstream regrind and cleaning capacity will be needed.

When the molybdenum head grade warrants operating the molybdenum plant, the bulk concentrate will be thickened in a conventional thickener (with no flocculant) and pumped to a differential flotation plant, where copper minerals will be depressed, and molybdenite floated into a molybdenum concentrate. The molybdenum concentrate will be filtered, dried, and packaged in tote bags for shipment to offshore roasters. Tailings from the molybdenum flotation circuit will constitute the copper concentrate, which will be pumped approximately 30 km to a filter plant at a port on the Caribbean coast. If the molybdenum head grade is very low, the molybdenum separation plant will be bypassed.




17.0 MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES

17.1 Mineral Resource Estimates

17.1.1 Introduction

Mineral resource estimates have been generated for all porphyry copper-type deposits that have been identified on the Cobre de Panamá concession and that have been drilled sufficiently to qualify as Inferred resources. These include the Botija, Colina, Valle Grande, Botija Abajo, and Brazo mineralized zones. The mineral resource estimates were prepared under the direction of Robert Sim, P.Geo, with the assistance of Bruce Davis, F.AusIMM. Both Mr. Sim and Dr. Davis meet the requirements of an independent Qualified Person under the standards of NI 43-101 as set out in Section 22 and the attached Certificate of Qualified Person.

The estimates are developed from 3D block models based on geostatistical applications using commercial mine planning software (MineSight® v4.60.09). The project limits area is based on the UTM coordinate system using the WGS84 ITRF-97 projection. A nominal block size of 25x25x15 m V is considered appropriate for the distribution of sample data, the deposit type, and the scale. Sample data are derived from surface diamond drill holes completed primarily from three main drilling programs beginning in the mid-late 1990s. Most of the drilling has been completed with vertical holes spaced on approximately 100 m intervals throughout the deposits. Drilling at Valle Grande is primarily completed with inclined drill holes designed to intersect what was originally interpreted to be a shallow-dipping mineralized target.

This report includes mineral resource estimates for the three better-defined porphyry deposits – Botija, Colina, and Valle Grande – as well as the lesser-drilled Botija-Abajo/ Brazo area. The locations of these deposit areas are shown in Figure 17-1. No mineral resource estimates are included in this report for the Molejón gold deposit, which is owned and operated by Petaquilla Minerals Ltd.

The resource estimate has been generated from drill hole sample assay results and the interpretation of a geologic model that relates to the spatial distribution of copper, molybdenum, gold, and silver. Interpolation characteristics have been defined based on the geology, drill hole spacing, and geostatistical analysis of the data. The resources have been classified by their proximity to the sample locations and are reported, as required by NI 43-101, according to the CIM standards on Mineral Resources and Reserves.




Figure 17-1: Drill Hole Plan Map

17.1.2 Geologic Model, Domains, and Coding

The Mina de Cobre property hosts a series of deposits containing appreciable amounts of copper and minor molybdenum, gold, and silver resulting from mineralization related to the intrusion of porphyritic rocks into pre-existing host rocks of andesitic and granodioritic composition. The larger deposits, Botija, Colina, and Valle Grande, and smaller Botija-Abajo and Brazo all occur over an area measuring approximately 10 km (E-W) x 4 km (N-S). Three-dimensional interpretations of the distribution of the rock types at Botija, Colina, and Valle Grande have been completed using drill hole and mapping information. The distributions of these rock type domains are shown in Figure 17-2, 17-3 and 17-4.

The current understanding of the geology at Botija-Abaja/Brazo does not allow for the generation of a 3D geology model at this stage; relogging of all drill core in this area is recommended and is underway.

Mineralization tends to occur within the porphyritic rocks and extending into the contact area of the plutonic rocks. There is evidence of minor post-mineral (barren) dikes, but these are often less than 3 m in width and, as a result, interpretation and segregation in the resource model are impractical at the overall scale of these deposits.

Copper mineralization consists primarily of chalcopyrite and minor bornite in hypogene mineralization. A poorly developed mixed zone of oxide copper minerals and secondary copper minerals such as chalcocite and covellite exists in the saprock zone overlying the hypogene mineralization. This material makes up a minor component of overall tonnage and is modelled separately from the hypogene mineralization.




Convincing structural features that influence the distribution of the mineralization have not been recognized. There has been no clear interpretation of either feeder zones or post-depositional zones of displacement. As a result, structural features have not been a factor in generating the resource model.

The geology database also includes alteration types in drill holes. Based on a review of this information, it shows a somewhat erratic distribution of facies and as such does not appear to be useful in controlling grade distribution. As a result a wireframe interpretation of the various alteration assemblages has not been undertaken. The grade properties of the various facies have been reviewed and are presented in the data analysis section of this report (Section 17.1.5).

Figure 17-2: Lithology Model Isometric




Figure 17-3: Botija Lithology Model Isometric




Figure 17-4: Colina and Valle Grande Lithology Model Isometric




17.1.3 Available Data

The drill hole database comes from four sources: original drilling done by PMRD (Japanese consortium) between 1970 and 1976, drilling conducted by Teck/Adrian in the early-mid 1990s, holes completed by the Petaquilla Copper Corp. (PTC) between 2006 and 2008 and, finally, drilling by MPSA in 2007-2009 (Figure 17-1).

Except for the Japanese holes, which contain only copper and molybdenum results (not tested for gold or silver), samples have been analyzed for copper, gold, molybdenum, and silver. Resource models were initially generated for all areas in August 2009 using drilling data dated up to 28 July 2009. During late summer and fall of 2009, additional infill drilling was completed on a portion of the Botija deposit, and the Botija model was subsequently updated using data dated 24 November 2009. Drilling was underway in the Brazo area in the fall of 2009, but the information was not included in the 24 November 2009 database. No additional drilling has been conducted on the Colina or Valle Grande deposits after the initial August 2009 model build.

Assay data were provided in the form of a series of Excel spreadsheet files, which were reformatted before being imported into MineSight®. Values below detection limits were handled as follows:

Japanese holes – If Cu value exists, set all missing Mo values to zero (0). If Mo exists, set all missing Cu values to zero (0). Au and Ag never tested; therefore, these all remain as “missing” (“-1” value in MineSight® treated as “missing”).

Teck/Adrian holes – If Cu value exists, set all missing Mo, Au, and Ag values to zero (0). If Mo exists, set all missing Cu values to zero (0).

MPSA holes – Values below DL are identified with “<#” in the spreadsheet file. These have been changed to 1/2DL before being imported into MineSight®. Note that original molybdenum analysis in MPSA holes are in Moppm units, which have been converted to Mo% (Mo%=Moppm/10000) to be consistent with pre-MPSA assay data.

Additional information includes logged geology data, including lithology and alteration codes. Bulk density data are also available for all MPSA drill holes. The surface topographic data are derived from Lidar information (April 2009), which has been regularized to a 10 m grid and triangulated into a 3D surface. This simplified the data for easier use, and the level of detail is considered sufficient for the scale of the deposits.

As of the August 2009 model build, the database has a total of 1,275 diamond core drill holes with a cumulative length of 230,555 m. This includes exploration and condemnation holes proximal to the main porphyry deposits plus all drilling conducted at the neighbouring Molejón deposit. It also includes several holes, drilled within the mineralized areas of the deposits, that were completed primarily to provide material for metallurgical testing. Some of these “Met” holes were not sampled and assayed, but they do provide information in the geologic interpretation of the deposits. During 2007 and 2008, Petaquilla Copper Limited (a shareholder




in MPSA at the time) drilled a number of relatively short holes in all deposit areas to test for the presence of oxide copper mineralization. These holes do not influence the resource estimation but do provide some definition of the saprolite, saprock, and hypogene mineralization limits.

The distribution of drilling data is summarized by area in Table 17-1. Drill holes are typically spaced at 100 m to140 m intervals throughout the copper porphyry deposit areas. In mid-2009 a portion of the central, higher-grade area of the Botija deposit was drilled with holes on a 70 m grid pattern.

Table 17-1: Distribution of Drilling Data by Area

Area No. of Drill Holes Total Length

(m) Comments

Botija 296 71,073 Assay sample data for 262 holes

Colina 233 52,734 Assay sample data for 202 holes

Valle Grande 242 43,560 Assay sample data for 233 holes

Botija-Abajo/Brazo 227 28,591 Assay sample data for 156 holes. No assay data for many PTC holes at BA

Molejón 130 15,341 Gold analysis only

Other 147 19,448 Exploration and condemnation drilling

Total 1,275 230,555

Individual sample intervals range from 0.03 m to 36.50 m and average 1.54 m in length; 86% of sample intervals are exactly 1.5 m in length. Most of the long sample intervals are derived from saprolite composites. The basic statistical summary of the assay sample database is provided in Table 17-2.

Core recovery data from MPSA and Teck-Adrian drill holes located in the Botija and Colina deposit areas were reviewed. Values range from 1% to 100% with an overall average of 93%. Less than 5% of the sample intervals have recoveries below 50%, and 90% of the data have recoveries greater than 80%. There are no indications of any correlation between metal grade and recovery. No data adjustments or exclusions related to core recoveries were carried out before the resource model was developed.

Figure 17-1 shows the distribution of drilling by vintage in relation to the various deposit areas.




Table 17-2: Statistical Summary of Sample Assay Data by Area

Area/Element No. of

Samples1 Total Length

(m) Min Max Mean2 Std Dev

Botija

Copper (%) 42,086 65,852 0 6.86 0.34 0.328

Molybdenum (%) 41,829 65,395 0 0.637 0.007 0.014

Gold (g/t) 39,495 60,068 0 2.33 0.07 0.099

Silver (g/t) 39,502 60,078 0 39.7 1.09 1.21

Colina

Copper (%) 30,991 48,960 0 8.07 0.33 0.312

Molybdenum (%) 30,990 48,932 0 0.64 0.006 0.012

Gold (g/t) 27,375 41,485 0 16.62 0.059 0.146

Silver (g/t) 27,380 41,492 0 690 1.4 5.2

Valle Grande

Copper (%) 26,722 41,356 0 12.81 0.29 0.372

Molybdenum (%) 26,688 41,308 0 0.537 0.006 0.012

Gold (g/t) 26,464 40,624 0 14.76 0.041 0.130

Silver (g/t) 26,480 40,650 0 209 1.2 2.03

Botija-Abajo/Brazo

Copper (%) 15,809 22,671 0 8.16 0.20 0.280

Molybdenum (%) 12,333 17,767 0 0.387 0.004 0.009

Gold (g/t) 15,809 22,671 0 40 0.086 0.407

Silver (g/t) 15,849 22,731 0 219 0.8 3.24

Note: 1 A small number of sample intervals were split at geology contacts when the data are loaded into MineSight®. Therefore, the total number of samples listed may be higher than the original data provided by MPSA.

2 Statistics are weighted by sample length.

17.1.4 Compositing

Drill hole samples are composited to standardize the database for further statistical evaluation. This step eliminates any effects related to the sample length that may exist in the data.

To retain the original characteristics of the underlying data, a composite length is selected that reasonably reflects the average original sample length. The generation of longer composites results in some degree of smoothing, which could mask certain features of the data. Sample intervals are relatively consistent in the database with approximately 86% at exactly 1.5 m in length. As a result, a standard composite length of 1.5 m has been applied to the sample data.

Drill hole composites are length-weighted and have been generated “down-the-hole,” meaning that composites begin at the top of each hole and are generated at 1.5 m intervals down the length of the hole. Several holes were randomly selected, and the composited values were checked for accuracy. No errors were found.




17.1.5 Exploratory Data Analysis

Exploratory data analysis (EDA) involves the statistical summarization of the database in order to better understand the characteristics of the data that may control grade. One of the main purposes of this exercise is to determine if there is evidence of spatial distinctions in grade that may require the separation and isolation of domains during interpolation. The application of separate domains prevents unwanted mixing of data during interpolation and the resulting grade model will better reflect the unique properties of the deposit. However, applying domain boundaries in areas where the data are not statistically unique may impose a bias in the distribution of grades in the model.

A domain boundary, which segregates the data during interpolation, is typically applied if the average grade in one domain is significantly different from that of another domain. A boundary may also be applied where there is evidence that there is a significant change in the grade distribution across the contact.

EDA was initially conducted for copper, molybdenum, gold, and silver between lithology and alteration types.

17.1.5.1 Comparison of Data by Vintage

The data generated by the various drilling campaigns were compared to ensure the absence of any potential bias that could affect resource estimations. The drilling data were segregated into three groups: MPSA (FEED) data drilled between 2007 and 2009, Teck/Adrian (T/A) data generated between 1992 and 1997, and combined Japanese (1970-1976) and PTC (2006-2008) (J+PTC); this information was combined because relatively few deep PTC drill holes contribute to the resource model development.

Drill hole data were composited to 15 m intervals and declustered through the development of a series of nearest-neighbour (NN) models. A 15 m composite length was selected because it equals the vertical size of the blocks in the model, ensuring that the analysis uses all of the data. “Distance to data” values were stored in blocks during the generation of all three NN models, allowing for subsequent comparisons to be limited to proximal data. Declustered blocks selected for comparisons were limited to a maximum distance of 100 m from both data types (i.e., comparisons between MPSA and Teck/Adrian data are limited to declustered blocks within a maximum distance of 100 m from both data types.) A series of QQ plots were developed for comparison purposes.

Overall, copper grades tend to agree in all deposit areas. Molybdenum grades in the new (FEED) holes show consistently lower grades, but because it is a relatively small economic contributor in the deposits, this is not considered to be a significant issue. Gold and silver tend to show similar grades between FEED and T/A holes, which tend to be higher than J+PTC holes. This may be the result of improved sampling and assaying procedures.

When present, the grade differences between data sources are considered relatively minor, indicating no significant bias between data sets derived from differing sources (vintages). This




comparison validates, by association, some of the older data and shows that all data can be combined for use in the development of resource models.

17.1.5.2 Basic Statistics by Domain

Summary statistics are evaluated using a series of boxplots for copper, molybdenum, gold, and silver based on the logged rock type and alteration facies designations. Based on the rock types assigned in the geologic logs:

Elevated copper grades are generally present in all three of the main rock types, andesite, porphyry, and granodiorite, and also in the minor breccia and faults units. The porphyry shows marginally higher grades at Botija and Colina. Low grades occur in the saprolite and post-mineral dikes.

Gold grades are slightly higher in the saprolite compared to the other rock domains. Molybdenum tends to be similar across all rock types. Silver trends are similar to copper.

Similar results are evident in boxplots generated using rock type codes derived from the 3D interpretation of lithologic types. These results indicate that the interpreted wireframe model is an appropriate representation of the underlying drill hole data. Examples of the copper distribution by rock type are shown in Figure 17-5 to 17-7.

Alteration trends show slightly higher copper, gold, silver, and molybdenum grades in phyllic, potassic, and, often to a lesser extent, argillic facies in all deposit areas. These are all minor alteration types that tend to be somewhat erratically distributed throughout the deposit areas. In total, they comprise less than 20% of the contained sample data.




Figure 17-5: Botija Boxplot for Copper by Rock Type

Figure 17-6: Colina Boxplot for Copper by Rock Type




Figure 17-7: Valle Grande Boxplot for Copper by Rock Type

17.1.5.3 Contact Profiles

The nature of grade trends between two domains is evaluated using a contact profile that graphically displays the average grades at increasing distances from the contact boundary. Contact profiles that show a marked difference in grade across a domain boundary are an indication that the two data sets should be isolated during interpolation. Conversely, if there is a more gradual change in grade across a contact, the introduction of a “hard” boundary (i.e., segregation during interpolation) may result in much different trends in the grade model. In this case the change in grade between domains in the model is often more abrupt than the trends seen in the raw data. Finally, a flat contact profile indicates no grade changes across the boundary. In the case of a flat profile, “hard” or “soft” domain boundaries will produce similar results in the model.

A series of contact profiles were generated to evaluate the change in copper grades across prominent rock domain boundaries. None were generated for the minor elements molybdenum, gold, and silver because no significant grade differences are evident in the boxplots.

Copper grade differences tend to be flat or transitional across all logged rock type domains except the saprolite, which tends to be significantly lower than the neighbouring domains. The transition between saprock and underlying (fresh) rocks also tends to be transitional throughout, suggesting that little depletion and/or enrichment has taken place in this domain. At Botija, the andesite copper grades appear to be drop across the boundary with the granodiorite (Figure 17-8) and porphyry (Figure 17-9), suggesting that hard boundaries may be required. However, upon inspection of the sample grades, several low-grade areas of andesite exist, but more commonly the grades appear to be transitional between neighbouring domains.




Contact profiles produced between logged alteration types show transitional or no changes in copper grade at all contacts.

Figure 17-8: Botija Contact Profile for Copper between Andesite and Granodiorite




Figure 17-9: Botija Contact Profile for Copper between Andesite and Porphyry

17.1.5.4 Modelling Implications

The results of the EDA show similar trends in all deposit areas. The near-surface saprolitic layer shows significantly lower copper grades, which tend to change abruptly in relation to the underlying saprock and hypogene rock domains. This indicates that a “hard” boundary is required here during copper grade interpolations. Contact profiles at Botija suggest that andesite copper grades are distinct compared to neighbouring porphyry and granodiorite. However, visual inspection shows that this is only a local feature, which is supported by similarities exhibited in boxplots. It should be noted that the density of drilling at Botija is sufficient to generate an appropriate distribution of copper grades about the andesite contacts in the resource block.

There are no indications that alteration facies play a significant role in the distribution of metals in any of the deposit areas.




17.1.5.5 Conclusions

Table 17-3 summarize the elements to be estimated in the models and how the various domains are applied during grade interpolations. In the absence of an interpreted geologic model for Botija-Abajo/Brazo, no separate domains are used in this area during grade interpolations in the model.

Table 17-3: Summary of Interpolation Domains

Domains Element Saprolite Saprock+Andesite+Porphyry+Granodiorite

Copper Hard Hard Molybdenum Soft Soft Gold Soft Soft Silver Soft Soft

17.1.6 Bulk Density Data

Bulk density data are limited to samples collected by MPSA during its drilling program conducted between 2007 and 2009. A total of 9,798 rock samples were tested for bulk density, of which 121 were removed due to questionable or anomalous results (values below 2.0 t/m3 and above 4.0 t/m3), leaving a total of 9,677 valid measurements.

Bulk density was measured using the “wet” method in which pieces of core, averaging 10 to 15 cm in length, are weighed in air (Wa) and again while submerged in water (Ww). The volume (V) is determined as V=Wa-Ww (assuming 1 cm3 of water = 1 gram) and the density (D) is determined as D = Wa/V. Core is not sealed in wax or cellophane before weighing. Observations of representative drill core indicate that rocks are not porous, and therefore it is assumed that these measurements reflect the true bulk densities of the rocks.

Since bulk density data are restricted to only MPSA drilling, the spatial distribution of samples is not appropriate for modelling purposes. Therefore, average values were determined based on the rock type distribution and were assigned in the resource model. The results are listed in Table 17-4. Based on limited bulk density measurements for saprolite, it has been assigned a value of 1.5 t/m3.

Table 17-4: Summary of Bulk Density Values by Rock Type

Minzone Domain Botija (t/m3)

Colina (t/m3)

Valle Grande (t/m3)

Saprolite 1.50 1.50 1.50 Saprock 2.59 2.54 2.54 Andesite 2.71 2.70 2.70 Porphyry 2.61 2.62 2.61 Granodiorite 2.64 2.65 2.70




Bulk density measurements are available for only one MPSA hole drilled in the Brazo zone. In the absence of more data, an average value of 2.65 t/m3 is assumed in determining resource tonnage in the Botija-Abajo/Brazo area.

17.1.7 Evaluation of Outlier Grades

Histograms and probability plots of the distribution of all elements were reviewed in order to identify the existence of anomalous outlier grades in the composite database. In addition, the distribution of copper samples was further evaluated using a decile analysis. The decile analysis for copper suggests that top-cutting may not be required for any of the deposits; however, some erratic values exist at the high-grade range on log-probability plots.

Potential outlier samples were reviewed visually for their locations in relation to the surrounding data. It was decided that potential outlier samples would be controlled through the use of outlier limitations during block grade interpolation. Samples above a defined threshold are limited to a maximum distance of influence of 50 m during grade estimates. The limits and resulting effects on the models are listed in Table 17-5.

The percentage of metal lost in the model due to outlier restrictions is considered appropriate for all elements.

Table 17-5: Summary of Outlier Grade Controls

Deposit/Element Threshold % Metal Lost in Model 1

Botija Copper (%) 2.5 0.2 Molybdenum (%) 0.2 0 Gold (g/t) 1 0.2 Silver (g/t) 10 0.3

Colina Copper (%) 3.0 0.1 Molybdenum (%) 0.15 1.7 Gold (g/t) 1.5 0.8 Silver (g/t) 15 3.6

Valle Grande Copper (%) 4.0 0.6 Molybdenum (%) 0.15 1.7 Gold (g/t) 1.5 1.0 Silver (g/t) 10 1.3

Botija-Abajo / Brazo Copper (%) 2.5 0.1 Molybdenum (%) 0.10 0.1 Gold (g/t) 2.5 1.5 Silver (g/t) 30 1.3

Note: 1 Loss in metal determined in Measured and Indicated class blocks in the model.




17.1.8 Variography

The degree of spatial variability and continuity in a mineral deposit tends to depend on both the distance and direction between points of comparison. Typically, the variability between samples increases as the distance between samples also increases. If the variability is related to the direction of comparison, then the deposit is said to exhibit anisotropic tendencies, which can be summarized by an ellipse fitted to the ranges in the different directions. The semi-variogram is a common function used to measure the spatial variability within a deposit.

The components of the variogram include the nugget, the sill, and the range. Often, samples compared over very short distances – even samples compared from the same location – show some degree of variability. As a result, the curve of the variogram often begins at a point on the y-axis above the origin. This point is called the “nugget.” The nugget is a measure of not only the natural variability of the data over very short distances but also of the variability that can be introduced due to errors during sample collection, preparation, and assaying.

The amount of variability between samples typically increases as the distance between the samples increases. Eventually, the degree of variability between samples reaches a constant, maximum value. This is called the “sill,” and the distance between samples at which this occurs is referred to as the “range.”

The spatial evaluation of the data in this report has been conducted using a correlogram rather than the traditional variogram. The correlogram is normalized to the variance of the data and is less sensitive to outlier values, generally giving cleaner results.

Correlograms were generated for the distribution of copper, molybdenum, gold, and silver using the commercial software package Sage 2001© developed by Isaacs & Co. Multidirectional correlograms were generated for composited drill hole samples; the results are summarized in Table 17-6.

Table 17-6: Variogram Parameters

1st Structure 2nd Structure

Element Nugget S1 S2 Range

(m) AZ Dip Range

(m) AZ Dip

BOTIJA

Copper 0.175 0.363 0.462 139 286 11 950 109 -3

Spherical 104 17 6 470 18 -24

76 134 78 189 25 66

Molybdenum 0.450 0.269 0.281 111 305 10 288 339 -14

Spherical 79 31 -20 244 257 29

12 60 67 119 47 57

Gold 0.200 0.357 0.443 137 302 -34 934 109 -17

Spherical 72 45 -19 408 9 -29

32 339 50 235 45 55

Silver 0.300 0.422 0.278 94 58 7 736 94 -17

Spherical 53 339 -56 318 359 -15




1st Structure 2nd Structure

Element Nugget S1 S2 Range

(m) AZ Dip Range

(m) AZ Dip

13 323 33 209 49 67

COLINA

Copper 0.200 0.565 0.235 140 51 -10 1,602 285 -5

Spherical 81 141 2 440 198 29

43 39 80 182 5 60

Molybdenum 0.350 0.451 0.199 119 49 -6 1,225 296 -13

Spherical 110 138 8 388 213 28

16 354 80 175 3 59

Gold 0.300 0.433 0.267 107 78 14 945 297 -17

Spherical 74 345 12 372 29 -5

9 216 72 174 314 72

Silver 0.500 0.422 0.078 89 337 54 902 108 1

Spherical 63 95 19 148 199 38

1 196 29 75 17 52

VALLE GRANDE

Copper 0.175 0.625 0.200 64 324 11 743 305 -6

Spherical 40 53 -4 246 35 -4

25 125 78 122 335 83

Molybdenum 0.350 0.561 0.089 74 81 -7 949 312 12

Spherical 21 355 26 411 218 18

17 157 62 205 75 68

Gold 0.250 0.513 0.237 95 44 -12 527 148 19

Spherical 65 132 11 153 56 4

11 1 74 94 315 71

Silver 0.300 0.566 0.134 60 277 22 489 131 15

Spherical 28 133 63 161 45 -12

11 13 14 155 353 71

BOTIJA-ABAJO/BRAZO

Copper 0.150 0.489 0.361 50 30 -54 534 314 40

Spherical 44 305 3 338 197 28

12 37 35 280 84 37

Molybdenum 0.400 0.446 0.154 144 5 -3 357 214 8

Spherical 40 96 -19 203 123 5

9 86 71 153 1 80

Gold 0.400 0.439 0.161 46 338 -22 1,116 296 -10

Spherical 37 73 -13 324 24 11

2 13 64 137 249 75

Silver 0.500 0.416 0.084 56 103 -3 2,401 137 -4

Spherical 34 17 53 485 47 3

5 191 36 274 175 85

Note: Correlograms conducted on 1.5 m DH composite data.




17.1.9 Model Set-Up and Limits

Block models were initialized in MineSight®; the dimensions are defined in Table 17-7 and shown in Figure 17-10. The selection of a nominal block size of 25x25x15 m V is considered appropriate with respect to the drill hole distribution and the deposit type and scale. Given the close proximity of the Colina and Valle Grande deposits, a single model was used to estimate the resources across these two zones.

Table 17-7: Block Model Limits

Direction Minimum Maximum Block Size

(m) # Blocks

Botija

East 537000 540300 25 132

North 975800 978300 25 100

Elevation -495 300 15 53

Colina and Valle Grande

East 532000 537000 25 200

North 973800 978300 25 180

Elevation -360 390 15 50

Botija-Abajo / Brazo

East 538800 542500 25 148

North 973200 976100 25 116

Elevation -390 330 15 48




Figure 17-10: Block Model Limits

Blocks in the model were coded on a majority basis with the various rock codes. During this stage, blocks along a domain boundary are coded if >50% of the block occurs within the boundaries of that domain.

The proportion of blocks that occur below the topographic surfaces are also calculated and stored within the model as individual percentage items. These values are used as a weighting factor in determining the in-situ resources for the deposit.

17.1.10 Interpolation Parameters

The block model grades for the four modelled elements (Cu, Mo, Au, Ag) were estimated using ordinary kriging (OK). The vast majority of the economic potential of the deposits is related to the contained copper. For this reason, the model development has concentrated on copper, and the estimations were validated using the Hermitian (Herco) polynomial change-of-support model (also referred to as the Discrete Gaussian correction). This method is described in more detail in the next section.

The OK models were generated with a relatively limited number samples in order to match the change-of-support or (Herco) grade distribution. This approach reduces the amount of smoothing (averaging) in the model and, while there may be some uncertainty on a localized




scale, produces reliable estimations of the recoverable grade and tonnage for the overall deposit. All grade estimations use length-weighted composite drill hole sample data. The interpolation parameters are summarized by domain in Table 17-8.

Table 17-8: Interpolation Parameters

Interpolation Domain

Search Ellipse Range(m) # Composites

X Y Z Min/block Max/block Max/hole Other

Botija

Copper in Saprolite 500 500 100 5 30 10

Copper in all other domains

500 500 100 35 120 30

Molybdenum 500 500 100 35 150 30 1 DH per octant

Gold 500 500 100 35 120 30

Silver 500 500 100 20 60 15 1 DH per quadrant

Colina



500 500 100 35 150 30

1 DH per octant


Gold 500 500 100 35 150 30

Silver 500 500 100 35 150 30 1 DH per quadrant

Valle Grande



500 500 100 35 150 30

1 DH per octant


Gold 500 500 100 35 150 30

Silver 500 500 100 35 150 30 1 DH per octant

Batija-Abajo/Brazo

Copper 300 300 100 35 120 30 1 DH per octant


Gold 300 300 100 35 120 30 1 DH per octant

Silver 300 300 100 35 120 30 1 DH per octant

17.1.11 Validation

The results of the modelling process were validated through several methods. These include a thorough visual review of the model grades in relation to the underlying drill hole sample grades, comparisons with the change-of-support model, comparisons with other estimation methods, and grade distribution comparisons using swath plots.

Visual Inspection

The block model was visually inspected in detail in both section and plan to ensure the desired results following interpolation. This includes confirmation of the proper coding of blocks within




the various domains. The distribution of block grades was compared relative to the drill hole samples to ensure their proper representation in the model.

Model Checks for Change of Support

The relative degree of smoothing in the block model estimates were evaluated using the Discrete Gaussian or Hermitian polynomial change-of-support method (described by Journel and Huijbregts, Mining Geostatistics, 1978). With this method, the distribution of the hypothetical block grades can be directly compared to the estimated (OK) model through the use of pseudo-grade/tonnage curves. Adjustments are made to the block model interpolation parameters until an acceptable match is made with the Herco distribution. In general, the estimated model should be slightly higher in tonnage and slightly lower in grade when compared to the Herco distribution at the projected cutoff grade. These differences account for selectivity and other potential ore-handling issues that commonly occur during mining.

The Herco (Hermitian correction) distribution is derived from the declustered composite grades that were adjusted to account for the change in support as one goes from smaller drill hole composite samples to the large blocks in the model. The transformation results in a less-skewed distribution but with the same mean as the original declustered samples.

It is expected that selectivity during mining will be based primarily on copper grades and that the secondary elements molybdenum, gold, and silver will contribute only as minor accessories. For completeness, however, Herco plots were generated for all elements in all deposits. An appropriate degree of correlation is evident in all models. Examples from copper models are shown in Figure 17-11 through 17-14.

Comparison of Interpolation Methods

For comparison purposes, additional grade models were generated using both the inverse distance weighted (ID) and nearest-neighbour (NN) interpolation methods. The NN model was based on data composited to 15 m intervals. The results of these models are compared to the OK models at various cutoff grades in a series of grade/tonnage graphs. Examples from the copper models are shown in Figure 17-12. Overall, there is very good correlation between models. Reproduction of the model using different methods tends to increase the confidence in the overall resource.




Figure 17-11: Herco Plots for Copper




Figure 17-12: Grade/Tonnage Comparison of Copper Models

Swath Plots (Drift Analysis)

A swath plot is a graphical display of the grade distribution derived from a series of bands, or swaths, generated in several directions through the deposit. Grade variations from the OK model are compared using the swath plot to the distribution derived from the declustered (NN) grade model.

On a local scale, the NN model does not provide reliable estimations of grade but, on a much larger scale, it represents an unbiased estimation of the grade distribution based on the underlying data. Therefore, if the OK model is unbiased, the grade trends may show local fluctuations on a swath plot, but the overall trend should be similar to the NN distribution of grade.

Swath plots have been generated in three orthogonal directions for all elements in all models. Examples from copper models are shown in Figure 17-13.




Figure 17-13: Swath Plots for Copper




There is good correspondence between the models in all elements. The degree of smoothing in the OK model is evident in the peaks and valleys shown in the swath plots. Deviations tend to occur for two reasons. First, reduced tonnages near the edges of the deposit tend to accentuate the differences in grade between models. Second, differences in grade become more apparent in the lower-grade areas – typically the flanks of the deposit where the drilling density often decreases.

17.1.12 Resource Classification

The mineral resources for the various deposits have been classified in accordance with the CIM definition standards for mineral resources and mineral reserves (CIM, 2005). The classification parameters are defined in relation to the distance to sample data and are intended to encompass zones of reasonably continuous mineralization.

The parameters are based on the results of a geostatistical study of uncertainty that defines categories based on confidence limits. Measured resources are defined as material in which the predicted grade is within ±15% on a quarterly basis, at a 90% confidence limit. In other words, there is a 90% chance that the recovered metal for a quarter-year of production will be within ±15% of the production actually achieved. Similarly, Indicated resources include material in which the yearly metal production is estimated with ±15% at the 90% confidence level.

The confidence limit evaluation shows similar results for the Botija and Colina deposits, with quarterly production estimated within ±15% at the 90% confidence limit with holes spaced at 70 m intervals. Annual production can be forecast at similar confidence levels based on drilling spaced at 150 m intervals.

At Valle Grande, the relative variogram shows more variation, and the classification criteria have been adjusted accordingly. Measured resources require drilling on a nominal 50 m grid pattern, and Indicated resources require drilling on a 120 m grid pattern.

The Botija-Abajo / Brazo area has not undergone a detailed evaluation of this type. The parameters used for Botija and Colina are assumed to be applicable for classification of resources in this area.

These results are used to define the classification criteria listed below. The limits of Indicated and Inferred resources are shown in plan view in Figure 17-14.

Measured Resources – Botija, Colina and Botija-Abajo / Brazo areas require model blocks with copper grades estimated by a minimum of three drill holes located within an average distance of 50 m. This is achieved with drill holes at a nominal spacing of 70 m. At Valle Grande, Measured model blocks are estimated by a minimum of three drill holes within an average distance of 35 m.




Indicated Resources – Botija, Colina, and Botija-Abajo / Brazo areas require model blocks with copper grades estimated by a minimum of three drill holes located within a maximum average distance of 100 m. This is achieved with drill holes at a nominal spacing of 150 m. At Valle Grande, Indicated model blocks are estimated using a minimum of three drill holes within an average distance of 90 m.

Inferred Resources – These are model blocks that do not meet the criteria for Measured or Indicated resources but are within a maximum distance of 250 m from a single drill hole.

Figure 17-14: Plan Limit of Resource Classes




17.1.1 Mineral Resource Estimates

Mineral resources are summarized at a series of copper cutoff grades in Table 17-9 to 17-12. The base case cutoff grade of 0.15% Cu is considered to be appropriate based on assumptions derived from deposits of similar type, scale, and location.

To ensure that the reported resource exhibits reasonable prospects for economic extraction, the resource is limited within a pit shell with a 40° slope angle generated about copper grades in blocks classified in the Indicated and Inferred categories at a copper price of $2.30/lb copper and total operating costs of $7.41/t. This constraint excludes deeper mineralization that cannot support the increased strip ratios at this assumed metal price. The following tables list the mineral resource estimates. Mineral resources, that are not mineral reserves, do not have demonstrated economic viability. It cannot be assumed that all or any part of an Inferred mineral resource will be upgraded to an Indicated or Measured mineral resource as a result of continued exploration.

There are no known factors related to environmental, permitting, legal, title, taxation, socioeconomics, marketing, or political issues that could materially affect the mineral resource.

Table 17-9: Summary of Measured Mineral Resources

Cutoff Grade (Cu%) Mt Cu % Mo % Au g/t Ag g/t

Botija

0.10 271 0.55 0.009 0.13 1.5

0.15 261 0.56 0.009 0.13 1.5

0.20 246 0.59 0.010 0.14 1.6

0.25 232 0.61 0.010 0.14 1.6

0.30 220 0.63 0.010 0.15 1.7

0.35 207 0.65 0.011 0.15 1.7

0.40 193 0.67 0.011 0.16 1.7

Note: There are no Measured resources at Colina, Valle Grande, Botija-Abajo, or Brazo. Mineral resources do not have demonstrated economic viability. The “base case” cutoff grade of 0.15%Cu is highlighted in the table.

Table 17-10: Summary of Indicated Mineral Resources


Botija

0.10 1,066 0.30 0.006 0.05 1.0

0.15 907 0.33 0.007 0.06 1.0

0.20 734 0.36 0.007 0.06 1.1

0.25 561 0.41 0.008 0.07 1.1

0.30 420 0.45 0.009 0.08 1.2

0.35 309 0.50 0.010 0.08 1.2

0.40 226 0.55 0.011 0.09 1.3





Colina

0.10 1,272 0.34 0.006 0.05 1.4

0.15 1,178 0.35 0.007 0.05 1.5

0.20 1,006 0.38 0.007 0.06 1.5

0.25 841 0.42 0.008 0.06 1.6

0.30 680 0.45 0.008 0.07 1.7

0.35 522 0.49 0.009 0.08 1.7

0.40 391 0.53 0.009 0.08 1.8

Valle Grande

0.10 717 0.33 0.006 0.04 1.3

0.15 671 0.34 0.006 0.04 1.3

0.20 591 0.37 0.007 0.04 1.4

0.25 494 0.40 0.007 0.05 1.5

0.30 390 0.43 0.007 0.05 1.6

0.35 289 0.47 0.008 0.06 1.7

0.40 200 0.51 0.008 0.06 1.8

Botija-Abajo

0.10 215 0.25 0.003 0.09 0.9

0.15 184 0.28 0.004 0.09 0.9

0.20 142 0.31 0.004 0.10 0.9

0.25 100 0.34 0.004 0.11 0.9

0.30 64 0.39 0.004 0.12 0.9

0.35 40 0.43 0.004 0.13 1.0

0.40 21 0.48 0.004 0.13 1.0

Brazo

0.10 74 0.41 0.004 0.12 0.7

0.15 71 0.43 0.004 0.12 0.7

0.20 66 0.45 0.004 0.12 0.7

0.25 61 0.46 0.004 0.12 0.8

0.30 56 0.48 0.004 0.13 0.7

0.35 51 0.50 0.004 0.13 0.7

0.40 44 0.52 0.004 0.13 0.7

All Areas Combined

0.10 3,343 0.32 0.006 0.05 1.2

0.15 3,010 0.34 0.006 0.06 1.2

0.20 2,539 0.37 0.007 0.06 1.3

0.25 2,057 0.41 0.007 0.07 1.4

0.30 1,611 0.44 0.008 0.07 1.5

0.35 1,211 0.49 0.008 0.08 1.5

0.40 883 0.53 0.009 0.08 1.6

Note: Mineral resources do not have demonstrated economic viability. The “base case” cutoff grade of 0.15% Cu is highlighted in the table.




Table 17-11: Summary of Measured And Indicated Mineral Resources

Cutoff Grade (Cu%)

Mt Cu % Mo % Au g/t Ag g/t

Botija

0.10 1,337 0.35 0.007 0.07 1.1

0.15 1,168 0.38 0.007 0.07 1.1

0.20 980 0.42 0.008 0.08 1.2

0.25 794 0.47 0.009 0.09 1.3

0.30 640 0.51 0.010 0.10 1.4

0.35 516 0.56 0.010 0.11 1.4

0.40 419 0.60 0.011 0.12 1.5

Colina

0.10 1,272 0.34 0.006 0.05 1.4

0.15 1,178 0.35 0.007 0.05 1.5

0.20 1,006 0.38 0.007 0.06 1.5

0.25 841 0.42 0.008 0.06 1.6

0.30 680 0.45 0.008 0.07 1.7

0.35 522 0.49 0.009 0.08 1.7

0.40 391 0.53 0.009 0.08 1.8

Valle Grande

0.10 717 0.33 0.006 0.04 1.3

0.15 671 0.34 0.006 0.04 1.3

0.20 591 0.37 0.007 0.04 1.4

0.25 494 0.40 0.007 0.05 1.5

0.30 390 0.43 0.007 0.05 1.6

0.35 289 0.47 0.008 0.06 1.7

0.40 200 0.51 0.008 0.06 1.8

Botija-Abajo

0.10 215 0.25 0.003 0.09 0.9

0.15 184 0.28 0.004 0.09 0.9

0.20 142 0.31 0.004 0.10 0.9

0.25 100 0.34 0.004 0.11 0.9

0.30 64 0.39 0.004 0.12 0.9

0.35 40 0.43 0.004 0.13 1.0

0.40 21 0.48 0.004 0.13 1.0

Brazo

0.10 74 0.41 0.004 0.12 0.7

0.15 71 0.43 0.004 0.12 0.7

0.20 66 0.45 0.004 0.12 0.7

0.25 61 0.46 0.004 0.12 0.8

0.30 56 0.48 0.004 0.13 0.7

0.35 51 0.50 0.004 0.13 0.7

0.40 44 0.52 0.004 0.13 0.7

All Areas Combined

0.10 3,615 0.34 0.006 0.06 1.2

0.15 3,271 0.36 0.007 0.06 1.3

0.20 2,785 0.39 0.007 0.07 1.3

0.25 2,290 0.43 0.008 0.07 1.4

0.30 1,831 0.47 0.008 0.08 1.5

0.35 1,418 0.51 0.009 0.09 1.5

0.40 1,076 0.55 0.009 0.10 1.6




Table 17-12: Summary of Inferred Mineral Resource


Botija 0.10 669 0.17 0.003 0.03 0.7 0.15 407 0.21 0.004 0.03 0.7 0.20 192 0.25 0.005 0.04 0.8 0.25 82 0.30 0.006 0.04 0.9 0.30 31 0.35 0.008 0.05 1.0 0.35 12 0.42 0.010 0.07 1.1 0.40 6 0.47 0.011 0.08 1.2 Colina 0.10 1,329 0.22 0.005 0.03 1.2 0.15 1,090 0.24 0.005 0.03 1.2 0.20 693 0.28 0.005 0.04 1.3 0.25 374 0.33 0.006 0.05 1.4 0.30 200 0.38 0.006 0.05 1.5 0.35 107 0.43 0.007 0.06 1.5 0.40 57 0.49 0.007 0.06 1.6 Valle Grande 0.1 1,341 0.22 0.004 0.03 1.0 0.15 1,141 0.24 0.005 0.03 1.0 0.20 784 0.27 0.005 0.03 1.1 0.25 438 0.31 0.006 0.04 1.3 0.30 212 0.36 0.006 0.04 1.4 0.35 93 0.40 0.006 0.04 1.6 0.40 42 0.45 0.007 0.04 1.7 Botija-Abajo 0.10 418 0.19 0.004 0.06 0.8 0.15 287 0.22 0.005 0.07 0.9 0.20 155 0.26 0.005 0.08 1.0 0.25 70 0.30 0.007 0.09 1.1 0.30 28 0.35 0.009 0.10 1.3 0.35 11 0.40 0.013 0.13 1.5 0.40 4 0.46 0.018 0.15 1.8

Brazo 0.10 320 0.25 0.004 0.06 0.5 0.15 269 0.27 0.004 0.07 0.6 0.20 195 0.31 0.004 0.08 0.6 0.25 135 0.35 0.005 0.09 0.7 0.30 91 0.38 0.005 0.10 0.7 0.35 59 0.42 0.005 0.10 0.7 0.40 34 0.45 0.005 0.11 0.8

All Areas Combined 0.10 4,076 0.21 0.004 0.04 0.9 0.15 3,194 0.24 0.005 0.04 1.0 0.20 2,019 0.27 0.005 0.04 1.1 0.25 1,099 0.32 0.006 0.05 1.2 0.30 562 0.37 0.006 0.06 1.3 0.35 282 0.42 0.007 0.06 1.4 0.40 143 0.47 0.007 0.07 1.4

Note: Mineral resources do not have demonstrated economic viability. The “base case” cutoff grade of 0.15% Cu is highlighted in the table.




17.1.2 Comparison with Pevious Estimate

The previous resource models for Botija, Colina, and Valle Grande were produced as the basis of a feasibility study conducted by H.A Simons in 1998 (Table 17-13). This study used lower copper prices, a lower milling rate, and an effective copper cutoff grade of 0.30% Cu.

Table 17-13: Historical Resource Estimate – 14 January 1998

Deposit Mt Cu % Mo ppm Au g/t

Measured Botija 379 0.56 105 0.113 Colina (formerly Petaquilla) 147 0.49 90 0.075 Valle Grande - - - - Indicated Botija 155 0.49 102 0.075 Colina (formerly Petaquilla) 269 0.46 90 0.075 Valle Grande - - - - Measured and Indicated Botija 534 0.54 104 0.104 Colina (formerly Petaquilla) 416 0.47 90 0.075 Valle Grande - - - - Total 950 0.51 98 0.091

Note: Resource estimated using a net smelter revenue cutoff of US$3.50/t based on copper price of $1.10/lb, gold price of $375/oz, molybdenum price of $3.50/lb, and recoveries of 86% for copper, 58% for gold, and 62% for molybdenum.

AMEC reviewed the resource models for Botija and Colina in the fall of 2007 (AMEC, 2007) to evaluate whether they were compliant with the CIM Definition Standards for Mineral Resources and Mineral Reserves (2005) as required by NI 43-101. The underlying data and information for this review were provided by Teck from the archived 1998 feasibility study (Simons, 1998). The parts of the historic database related to the Valle Grande deposit were not located during this evaluation, and so AMEC’s review included only the Colina and Botija deposits.

During its evaluation of the underlying database (AMEC, 2007), AMEC found that the Japanese drill holes lacked documentation but that, based on comparisons with surrounding current (and validated) drilling, this older information was usable for resource estimation purposes. AMEC considered the QA/QC procedures used by Adrian and Teck in the 1990s to be inconsistent with current standards; however, the information was considered to have been validated through a combination of check assays and twin hole comparisons. This conclusion has been confirmed further by additional drill data generated by MPSA.

The Simons 1998 report states the following with regard to classification: “Classification turned out to be a problem. The original Teck model was classified by kriging variance which could not be duplicated as per the documentation. It was decided that a new classification scheme be adopted based on distance from samples.” Unfortunately, the classification parameters are not listed in the AMEC 2007 report but have been interpreted based on information presented in Appendix 2H of the Simons 1998 report:




Measured Resources (1998) – Blocks with a minimum of two drill holes within a 100x100x100 m search and within a maximum distance of 35 m from a single hole.

Indicated Resources (1998) – Blocks with a minimum of two holes within 100 m search and within 75 m from a single hole. Or, blocks that do not meet the criteria for measured resources but are within a maximum distance of 35 m from a single drill hole.

These parameters allow for the classification of both Measured and, to a lesser extent, Indicated resources based on a single drill hole, resulting in a “spotted” or clustered distribution of Measured resources about individual drill holes. The FEED Study interprets that in order to be consistent with the classification descriptions outlined by the CIM (2005) and required by NI 43-101, the Measured resources must be contiguous. As a result, Measured resources have been reduced from earlier estimates.

Except for the starter pit zone at Botija, most of the drilling in the deposit areas is spaced at intervals of 140 m or less. In this study, this sampling configuration results in the designation of much of the resource as Indicated. In the Botija starter pit area, drilling is at tighter spacing, and it qualifies as Measured resource.

17.2 Mine Planning & Mineral Reserves

17.2.1 Pit Limit Analyses

The mine plans were developed from the deposit models described earlier in this section. A series of floating cone analyses were conducted to determine economic pit limits and the mining phase development sequence for three mineral deposits in the project Concession area: Botija, Colina, and Valle Grande.

17.2.2 Recoveries

Only sulphide mineralization classified as Measured or Indicated was considered potential ore in saprock, andesite, porphyry, and granodiorite rock types; all saprolite and mineral resources classified as Inferred were treated as waste.

Copper recovery estimates are a function of the Cu grade expressed in the following formula:

Cu Recovery (in %) = 7.4823 * Ln(Cu%) + 94.329

The copper recovery was reduced by four percentage points for all rock types in Valle Grande. Copper recoveries were capped at 95%.

Molybdenum recovery is a function of both Cu and Mo grades. Figure 17-15 shows the Mo recovery functions by grade range, both graphically and in tabular form. Mo recovery in Valle Grande was reduced by three percentage points from the computed values.




Gold recovery was computed from the following formula, with the gold grade expressed in grams per tonne:

Au Recovery (in %) = 10.997 * Ln(Ag g/t) + 81.990

Silver recoveries were fixed at 47.3% for all rock types except saprock. Saprock recoveries were based on Botija/Colina granodiorite formulas, but reduced by 25 percentage points for all metals due to the partially oxidized nature of the material. Cu, Mo, and Au recoveries for low-grade andesite, porphyry, and granodiorite ore that will be stockpiled (i.e., for project return optimization) were reduced by 20 percentage points to account for anticipated partial oxidation.

Over the life of the project, concentrator recoveries are forecast to average about 86% for Cu, 59% for Mo, 54% for Au, and 46% for Ag.




Figure 17-15: Molybdenum Recoveries by Grade




17.2.3 Overall Slope Angles

Overall slope angles used in the FC evaluation were derived from slope guidelines provided by AMEC EARTH & ENVIRONMENTAL (AMEC E&E). The maximum bench face angle in Colina Sector 4 was reduced from 70° to 65° to better fit excavations using electric rope shovels, effectively treating the small Sector 4 slope the same as neighbouring Sector 3 to the east. The overall slope angles were then adjusted to accommodate the recommended maximum inter-ramp slopes, diversion ditches, and sumps, and the anticipated placement of haulage ramps along certain pit walls. The resulting overall slopes used in the FC evaluations are summarized in Table 17-14.

Table 17-14: Overall Slopes for Floating Cone Analyses

Slope Sector Pit Wall Overall Angle

(degrees)

Botija 1 rock SE, S, NW 34

Botija 2 rock SW 35

Botija 3 rock W 37

Botija 4 rock N, NE, E 40

Colina 1 rock SE, SW 35

Colina 2 rock NW, Pit Bottom 37

Colina 3 rock N, E 40-41

Colina 3 rock S 35

Colina 4 rock SSW 35

Valle Grande rock All 35-37

Low RQD All -

Saprolite All -

17.2.4 Price Sensitivity Evaluations

Mining cost assumptions were $1.21/t for ore and $1.54/t for waste. (Waste hauls, on average, are expected to be 2 km longer or more). This yields a weighted average cost of approximately $1.33/t. Sustaining mine capital costs were estimated at $0.30/t. The base ore processing and G&A costs were projected at $3.88/t and $1.49/t, respectively, for a nominal grinding rate of 150,000 t/d. Ore processing and G&A costs were then adjusted for anticipated grinding rates that vary by rock type and deposit. Freight, smelting and refining (FSR) costs of $0.33/lb Cu and $0.09/lb Mo were deducted from the market metals prices. Recovery and net smelter return (NSR) values were computed for each block by custom subroutines and stored within the block model.

Price sensitivities were evaluated in generally 5%, or $0.10/lb Cu, increments, ranging from prices at +25% ($2.50 Cu) to -55% ($0.90 Cu) of the base case of $2.00/lb Cu, $12.00/lb Mo, $750/oz Au, and $12.50/oz Ag. Prices for Mo, Au, and Ag were varied proportionately in all cases. For purposes of the pit limit evaluations, all saprock was treated as waste. Time value of money effects were included in the FC analyses with a 1.5% per bench discounting factor




that approximates an annual discount rate of 8-10%. This reduces the net present value of lower ore zones with respect to overlying waste, which must be stripped two to four years in advance of ore mining. Table 17-15 summarizes the combined FC price sensitivity results for all three deposits, excluding ore-grade saprock mineral resources. The base case FC results are highlighted in bold type.

The estimates presented in Table 17-15 should not be confused with mineral reserves, which are based on open pit designs that incorporate access, operating, geotechnical and other criteria in addition to economic constraints. The floating cone results should not be relied upon, but do provide an indication of potential mineral reserves that must be validated by proper designs. Mineral resources that are not mineral reserves do not have demonstrated economic viability.




Table 17-15: Floating Cone Price Sensitivity Analyses (combined results from Botija, Colina and Valle Grande)

Metal Price Contained Mineral Resources* (>= Variable Internal Cutoffs) Waste kt

Total kt

Strip Ratio Cu $/lb Mo $/lb Au $/oz Ag $/oz kt NSR $/t Cu % Mo % Au g/t Ag g/t

2.50 15.00 938 15.63 2,637,174 16.93 0.37 0.007 0.06 1.32 1,310,961 3,948,135 0.50

2.40 14.40 900 15.00 2,563,867 16.39 0.38 0.007 0.06 1.34 1,276,017 3,839,884 0.50

2.30 13.80 863 14.38 2,477,956 15.84 0.38 0.007 0.06 1.35 1,210,737 3,688,693 0.49

2.20 13.20 825 13.75 2,390,619 15.28 0.39 0.007 0.07 1.36 1,162,885 3,553,504 0.49

2.10 12.60 788 13.13 2,283,037 14.74 0.39 0.007 0.07 1.37 1,082,393 3,365,430 0.47

2.00 12.00 750 12.50 2,187,930 14.17 0.40 0.007 0.07 1.38 1,036,278 3,224,208 0.47

1.90 11.40 713 11.88 2,062,835 13.62 0.41 0.007 0.07 1.39 953,372 3,016,207 0.46

1.80 10.80 675 11.25 1,932,878 13.07 0.42 0.008 0.07 1.41 884,969 2,817,847 0.46

1.70 10.20 638 10.63 1,803,905 12.50 0.43 0.008 0.07 1.42 840,097 2,644,002 0.47

1.60 9.60 600 10.00 1,643,889 11.97 0.44 0.008 0.08 1.44 768,508 2,412,397 0.47

1.50 9.00 563 9.38 1,491,085 11.41 0.45 0.008 0.08 1.46 727,731 2,218,816 0.49

1.40 8.40 525 8.75 1,299,115 10.89 0.47 0.009 0.08 1.49 655,625 1,954,740 0.50

1.30 7.80 488 8.13 1,086,452 10.37 0.49 0.009 0.09 1.52 574,489 1,660,941 0.53

1.20 7.20 450 7.50 827,144 9.96 0.52 0.009 0.09 1.55 476,179 1,303,323 0.58

1.15 6.90 431 7.19 705,363 9.75 0.53 0.009 0.10 1.58 424,295 1,129,658 0.60

1.10 6.60 413 6.88 485,511 9.48 0.55 0.010 0.10 1.63 255,232 740,743 0.53

1.05 6.30 394 6.56 385,095 9.26 0.57 0.010 0.10 1.64 211,602 596,697 0.55

1.00 6.00 375 6.25 293,664 8.98 0.59 0.010 0.10 1.67 164,140 457,804 0.56

0.90 5.40 338 5.63 106,395 8.57 0.65 0.013 0.11 1.69 62,765 169,160 0.59

* Measured and Indicated. Excludes saprolite, saprock and inferred mineral resources, which are treated as waste in this tabulation.




17.3 Mining Pit/Phase Designs

The ultimate pits and internal mining phases were designed to accommodate large-scale mining equipment operating on 15 m benches. This equipment includes rotary blasthole drills capable of drilling holes up to 311 mm in diameter, 55 m3 electric shovels, 38 m3 front-end loaders, and off-highway haulage trucks with payload capacities of 360 tonnes.

The base case floating cone shells, derived from metal prices of $2.00/lb Cu, $12.00/lb Mo, $750/oz Au, and $12.50/oz Ag, as described in the preceding section, were used to guide the ultimate pit designs. Pit walls were smoothed to minimize or eliminate, where possible, noses and notches that could affect slope stability. Internal haulage ramps were included in each pit/phase design to allow for truck access to working faces on each level. Provisions were also made for pit dewatering collection ditches and sumps on 90 m vertical intervals within each pit. The basic parameters used in the design of the mining phases, or pushbacks, are summarized in Table 17-16.

Table 17-16: Basic Pit Design Parameters

Parameters Unit Value

Bench height m 15

Haul road width (including ditch & safety berm) m 43

Internal ramp gradient (typical) % 10

Minimum pushback width m 75

Typical pushback width m 90+

Inter-ramp slope angles used in the ultimate pit and internal phase designs were derived from slope guidelines provided by AMEC E&E. Table 17-17 summarizes the pit slope design parameters, including internal working slopes. Eleven mining phases were designed for the project: four for the Botija pit, four for Colina, and three for Valle Grande (VG). This total excludes a construction materials quarry that will be used to supply low-sulphur waste rock for project facilities during the preproduction period. Figure 17-16 illustrates the ultimate pit and waste rock storage facility (WRSF) designs that were developed for this study. Grid lines are on 2 km intervals.




Table 17-17: Pit Slope Design Parameters

Slope Sector Pit Wall/ Sector

Bench Face Angle

(degrees)

Catch Bench Interval

(m)

Catch Bench Width

(m)

IR Angle within Stack

(degrees)

Stack Height

(m)

Step-out Width

(m)

IR Angle including Step-outs (degrees)

IR Angle including Drainage* (degrees)

Ultimate Botija 1 rock SE, S, NW 50 30 11 39.7 90 16.5 38.3 36.6 Botija 2 rock SW 55 30 11 43.1 90 16.5 41.5 39.5

Botija 3 rock W 60 30 11 46.6 90 16.5 44.8 42.5 Botija 4 rock N, NE, E 65 30 11 50.2 90 16.5 48.2 45.6

Colina 1 rock SE, SW 55 30 11 43.1 90 16.5 41.5 39.5 Colina 2 rock NW, Pit Bottom 60 30 11 46.6 90 16.5 44.8 42.5 Colina 3 rock N, E, S 65 30 11 50.2 90 16.5 48.2 45.6

Colina 4 rock SSW 65 30 11 50.2 90 16.5 48.2 45.6

VG rock All 55 30 11 43.1 90 16.5 41.5 39.5 Low RQD All 55 15 10 36.2 60 20.0 33.1 -

Saprolite All 55 10 8 33.7 30 20.0 27.8 -

Working

37.5 degree Bot 1, VG 55 15 9 37.5 - - - - 40 degree Bot 2, Col 1 60 15 9 40.0 - - - - 43 degree Bot 3, Col 2 65 15 9 43.0 - - - -

45 degree Bot 4, Col 3-4 65 15 8 45.0 - - - -

* 24 m drainage benches on 90 m vertical intervals




Figure 17-16: Ultimate Pit and Waste Rock Storage Facility Designs




17.4 Mineral Reserve Estimates

The mineral reserve estimates for the Mina de Cobre Panamá project are based on material classified as proven and probable only; all Inferred mineral resources were treated as waste. Proven and probable mineral reserves generally correspond to Measured and Indicated mineral resource classifications, respectively, once economic, technical, environmental, socio/political, and other criteria are applied to define ore.

17.4.1 Ore Definition Parameters

Sulphide ore within the open pit mining sequence is defined by metal prices of $2.00/lb Cu, $12.00/lb Mo, $750/oz Au, and $12.50/oz Ag, and by the recoveries outlined in Section 17.2.2. Table 17-18 summarizes the economic parameters used to define ore in the mineral reserve estimates. Net smelter return (NSR) values were calculated for each model block using these parameters for saprock, andesite, porphyry, and granodiorite rock types; all saprolite was treated as waste. NSR cutoff grades for ore types, excluding saprock, vary by rock type, deposit, and time period for the purpose of increasing the present value of the project returns. The NSR cutoffs by ore type and deposit are presented in Table 17-19.

Saprock will not be shipped directly to the mill because of its lower anticipated recoveries but will be stockpiled for later processing if found to be above an NSR cutoff of $5.21/t. Similarly, mineralized andesite, porphyry, and granodiorite below the declining cutoffs (Years 1-15), but above the low-grade cutoffs, will be placed into a low-grade stockpile. All ore stockpiles are scheduled to be rehandled after the open pits are depleted.




Table 17-18: Ore Definition Parameters for Mineral Reserve Estimates

Sulphide Ore

Granodiorite Andesite Porphyry Item Unit Botija/Colina VG Botija/Colina VG Botija/Colina VG

Direct to Mill Ore Mill throughput t/d 164,000 144,000 137,000 112,000 164,000 144,000 Cu recovery % 80.2 77.5 81.6 79.4 80.2 77.5 Mo recovery (thru Mo con) % 60.7 57.7 60.7 57.7 60.7 57.7 Au recovery % 52.7 52.7 52.7 52.7 52.7 52.7 Ag recovery % 47.3 47.3 47.3 47.3 47.3 47.3 Cu con – Cu grade % 27.9 27.9 27.9 27.9 27.9 27.9 Cu con ocean freight $/dmt 68.00 68.00 68.00 68.00 68.00 68.00 Cu con treatment) $/dmt 80.00 80.00 80.00 80.00 80.00 80.00 Cu refining $/lb payable 0.080 0.080 0.080 0.080 0.080 0.080 Au refining $/oz payable 5.50 5.50 5.50 5.50 5.50 5.50 Ag refining $/oz payable 0.40 0.40 0.40 0.40 0.40 0.40 Cu payable % 96.43 96.43 96.43 96.43 96.43 96.43 Au payable % 92.0 92.0 92.0 92.0 92.0 92.0 Ag payable % 90.0 90.0 90.0 90.0 90.0 90.0 Cu FSR $/lb Cu payable 0.330 0.330 0.330 0.330 0.330 0.330 Mo con – Mo grade % 52.0 52.0 52.0 52.0 52.0 52.0 Mo con ocean freight $/dmt 90.00 90.00 90.00 90.00 90.00 90.00 Mo payable % 86.2 86.2 86.2 86.2 86.2 86.2 Mo FSR $/lb Mo payable 0.091 0.091 0.091 0.091 0.091 0.091 Royalties (NSR basis) % 2.00 2.00 2.00 2.00 2.00 2.00 Ore mining $/t 1.21 1.21 1.21 1.21 1.21 1.21 Waste mining $/t 1.54 1.54 1.54 1.54 1.54 1.54 Sustaining mine capital $/t mined 0.30 0.30 0.30 0.30 0.30 0.30 Basis ore process $/t @ 150,000 t/d 3.88 3.88 3.88 3.88 3.88 3.88 Basis general/admin $/t @ 150,000 t/d 1.49 1.49 1.49 1.49 1.49 1.49 Throughput adjusted ore processing $/t 3.55 4.04 4.25 5.20 3.55 4.04 Throughput adjusted general/admin $/t 1.36 1.55 1.63 2.00 1.36 1.55 Internal cutoff NSR $/t 4.58 5.26 5.55 6.86 4.58 5.26 Breakeven cutoff NSR $/t 6.42 7.10 7.39 8.70 6.42 7.10

Low Grade Ore Stockpile LG stockpile rehandling cost $/t 0.63 0.63 0.63 0.63 0.63 0.63 Model Cu recovery at LG stockpile cutoff

% 83.1 83.9 84.3 85.6 83.1 83.9

Recovery losses (Cu, Mo & Au) % 20.0 20.0 20.0 20.0 20.0 20.0 Stockpile cutoff blk model NSR $/t 6.86 7.74 8.10 9.78 6.86 7.74

Saprock Ore Recovery deduction, all metals % 25.0 - - - - - Stockpile rehandling cost $/t 0.63 - - - - - Stockpile cutoff blk model NSR $/t 5.21 - - - - -




Table 17-19: NSR Cutoff Grades (US$/t) used in Mineral Reserve Estimates

Ore Type Deposit PP Y1 Y2 Y3 Y4 Y5 Y6

Andesite Botija & Colina 11.15 11.15 11.15 11.95 11.15 11.95 11.15

Valle Grande 12.46 12.46 12.46 13.26 12.46 13.26 12.46

Porphyry Botija & Colina 10.18 10.18 10.18 10.98 10.18 10.98 10.18

Valle Grande 10.86 10.86 10.86 11.66 10.86 11.66 10.86

Granodiorite Botija & Colina 10.18 10.18 10.18 10.98 10.18 10.98 10.98

Valle Grande 10.86 10.86 10.86 11.66 10.86 11.66 11.66

Ore Type Deposit Y7 Y8 Y9 Y10 Y11-Y15 Y16-Y30 Low Grade

Andesite Botija & Colina 11.15 10.35 9.55 8.75 6.35 5.55 8.10

Valle Grande 12.46 11.66 10.86 10.06 7.66 6.86 9.78

Porphyry Botija & Colina 10.18 9.38 8.58 7.78 5.38 4.58 6.86

Valle Grande 10.86 10.06 9.26 8.46 6.06 5.26 7.74

Granodiorite Botija & Colina 10.18 9.38 8.58 7.78 5.38 4.58 6.86

Valle Grande 10.86 10.06 9.26 8.46 6.06 5.26 7.74

17.4.2 Material Densities

Bulk material densities were stored in each block model and used to calculate tonnages for mineral reserve estimates. Table 17-20 presents the average material densities used in the mineral reserve estimates.

Table 17-20: Material Densities used in Mineral Reserve Estimates

Material Botija(t/m3)

Colina(t/m3)

Valle Grande (t/m3)

Saprolite 1.50 1.50 1.50

Saprock 2.59 2.54 2.54

Andesite 2.71 2.70 2.70

Porphyry 2.61 2.62 2.61

Granodiorite 2.64 2.65 2.70

Average rock (excl saprolite) 2.64 2.65 2.68

17.4.3 Dilution

The Botija, Colina, and Valle Grande deposits are all well-disseminated Cu-Mo-Au-Ag mineralized systems that have large ore zones above the anticipated cutoff grades. The sample compositing and block grade interpolation process used to construct the deposit block models is believed to incorporate sufficient dilution and, hence, no additional dilution factors were applied.




17.4.4 Ore Recovery

Ore recovery within the open pits is expected to be virtually 100%. However, stockpiled ore losses are estimated at 7% to account for inaccessibility along steep terrain and for material losses in the contact zone with felled timber.

17.4.5 Mineral Reserve Estimate Summary

Total proven and probable mineral reserves are estimated at approximately 2.14 Gt grading 0.41% Cu, 0.008% Mo, 0.07 Au g/t, and 1.43 Ag g/t. Of this total, about 245 Mt in the Botija pit are classified as proven mineral reserves at average grades of 0.59% Cu, 0.010% Mo, 0.14 Au g/t, and 1.61 Ag g/t. Approximately 1.97 Gt of the total will be shipped either directly to the primary crusher or to a nearby ROM ore stockpile. About 173 Mt will be recovered from long-term stockpiles of saprock and low-grade ore. Table 17-21 and Table 17-22 summarize the proven and probable mineral reserve estimates by ore type, respectively. Table 17-23 presents the combined total of estimated proven and probable mineral reserves by deposit and ore type.

Table 17-21: Proven Mineral Reserve Estimates by Ore Type

Proven Mineral Reserves

Pit / Ore Type kt NSR $/t Cu % Mo % Au g/t Ag g/t

Botija

Saprock 4,400 11.78 0.43 0.013 0.11 1.62

Andesite 19,500 16.67 0.45 0.010 0.09 1.05

Porphyry 117,100 24.26 0.64 0.010 0.16 1.74

Granodiorite 104,300 21.25 0.57 0.010 0.13 1.57

Total 245,300 22.15 0.59 0.010 0.14 1.61

Note: Estimates based on January 2010 Botija and November 2009 Colina-VG deposit models and metal prices of $2.00/lb Cu, $12.00/lb Mo, $750/oz Au, and $12.50/oz Ag




Table 17-22: Probable Mineral Reserve Estimates by Ore Type

Probable Mineral Reserves


Botija

Saprock 8,400 9.26 0.37 0.008 0.07 1.58

Andesite 32,700 13.64 0.38 0.008 0.07 0.94

Porphyry 148,500 15.77 0.43 0.009 0.09 1.22

Granodiorite 369,500 12.71 0.36 0.007 0.06 1.11

Total Botija 559,100 13.53 0.38 0.008 0.07 1.13

Colina

Saprock 31,800 10.01 0.39 0.007 0.11 1.45

Andesite 210,500 14.08 0.40 0.007 0.07 1.54

Porphyry 542,900 14.68 0.41 0.007 0.07 1.58

Granodiorite 89,700 11.44 0.32 0.008 0.04 1.44

Total Colina 874,800 14.03 0.40 0.007 0.07 1.55

Valle Grande

Saprock 15,900 9.06 0.39 0.008 0.07 1.55

Andesite 171,500 12.79 0.39 0.006 0.05 1.36

Porphyry 236,100 13.29 0.40 0.008 0.05 1.56

Granodiorite 39,900 13.50 0.41 0.008 0.04 1.51

Total VG 463,400 12.98 0.39 0.007 0.05 1.48

Total All Pits

Saprock 56,100 9.63 0.39 0.007 0.09 1.49

Andesite 414,700 13.52 0.39 0.007 0.06 1.42

Porphyry 927,500 14.50 0.41 0.007 0.07 1.52

Granodiorite 499,100 12.55 0.36 0.007 0.05 1.20

Total 1,897,400 13.63 0.39 0.007 0.06 1.41





Table 17-23: Combined Proven and Probable Mineral Reserve Estimates by Ore Type

Proven + Probable Mineral Reserves


Botija

Saprock 12,800 10.12 0.39 0.010 0.09 1.59

Andesite 52,200 14.77 0.41 0.009 0.08 0.98

Porphyry 265,500 19.51 0.52 0.009 0.12 1.45

Granodiorite 473,800 14.59 0.41 0.008 0.07 1.21

Total Botija 804,400 16.16 0.44 0.008 0.09 1.28

Colina

Saprock 31,800 10.01 0.39 0.007 0.11 1.45

Andesite 210,500 14.08 0.40 0.007 0.07 1.54

Porphyry 542,900 14.68 0.41 0.007 0.07 1.58

Granodiorite 89,700 11.44 0.32 0.008 0.04 1.44

Total Colina 874,800 14.03 0.40 0.007 0.07 1.55

Valle Grande

Saprock 15,900 9.06 0.39 0.008 0.07 1.55

Andesite 171,500 12.79 0.39 0.006 0.05 1.36

Porphyry 236,100 13.29 0.40 0.008 0.05 1.56

Granodiorite 39,900 13.50 0.41 0.008 0.04 1.51

Total VG 463,400 12.98 0.39 0.007 0.05 1.48

Total All Pits

Saprock 60,500 9.79 0.39 0.008 0.09 1.50

Andesite 434,200 13.66 0.39 0.007 0.06 1.40

Porphyry 1,044,500 15.59 0.44 0.008 0.08 1.54

Granodiorite 603,400 14.05 0.39 0.008 0.07 1.26

Total 2,142,600 14.60 0.41 0.008 0.07 1.43


Total material (ore, stockpiled ore, and waste) within the designed ultimate pits is estimated at 3.444 Gt, resulting in a stripping ratio of 0.61:1. This excludes the effects of ore stockpile rehandling.

Contained metal in the mineral reserve is projected at approximately 19.6 billion pounds of copper, 361 million pounds of molybdenum, 4.96 million troy ounces of gold, and 98.7 million troy ounces of silver. All mineral reserve estimates reported in the Table 17-21, Table 17-22 and Table 17-23 are contained within the mineral resource estimates presented in Section 17.1.




It should be noted that the designed ultimate pits contain approximately 260 Mt of ore-grade Inferred mineral resources that are treated as waste in the above estimates. These Inferred mineral resources have an average grade of 0.28% Cu, 0.005% Mo, 0.06 Au g/t, and 1.14 Ag g/t, and will require further drilling to potentially be considered as mineral reserves that could lower the average stripping ratio. Inferred mineral resources are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves. Inferred mineral resources have a great amount of uncertainty as to their existence and as to whether they can be mined legally or economically. It cannot be assumed that all or any part of inferred mineral resources will ever be upgraded.

Failure to obtain the necessary permits to develop the project could affect all of the above mineral reserve estimates. There are no other known factors related to environmental, permitting, legal, title, taxation, socio-economic, marketing, or political issues that could materially affect the mineral reserve estimates.

The mineral reserves presented in this section are as of March 31, 2010 and were estimated by William Rose, P.E., Principal Mining Engineer of WLR Consulting, Inc. Mr. Rose meets the requirements of an independent Qualified Person under the standards of NI 43-101 as set out in Section 22 and the attached Certificate of Qualified Person.




18.0 OTHER RELEVANT DATA AND INFORMATION

To the best of the Authors’ knowledge, all relevant data and information has been addressed elsewhere in this report.




19.0 INTERPRETATION AND CONCLUSIONS

The following conclusions are drawn from the completion of the FEED Study:

1. Copper-gold-molybdenum porphyry-style mineralization occurs at the southern margin of a large mid-Oligocene (36.4 Ma) granodioritic batholith located in north central Panamá. A 9 km by 4.5 km WNW-ESE oriented zone contains three large deposits: Botija, Colina and Valle Grande and a number of smaller zones, the most significant being Brazo and Botija Abajo.

2. Extensive drilling (1,275 holes totaling 230,555 m) has defined a mineral resource (measured and indicated) of 3,271 Mt grading 0.36% Cu, 0.007% Mo, 0.06 g/t Au and 1.30 g/t Ag. A mineral reserve (proven and probable) of 2,143 Mt grading 0.41 % Cu, 0.008 % Mo, 0.07 g/t Au and 1.43 g/t Ag is included in this resource estimate.

3. Proper diligence has been taken with respect to the drilling program, sampling and metallurgical testing – meeting or exceeding industry standard practices. The data density and reliability are sufficient to support a feasibility-grade evaluation of the project.

4. The estimated mineral reserves will support mining operations for approximately 30 years at nominal ore processing rates of 150,000 t/d through the first nine years of operation and 225,000 t/d thereafter. Ore processing will be scheduled for continuous operation and will consist of conventional crushing, grinding and flotation of sulphide mineralization. Metallurgical testing indicates that the concentrates will be of good quality, with no significant levels of deleterious constituents.

5. Significant infrastructure must be developed locally to support a project of this scale, including, but not limited to, the construction of port facilities, a concentrate pipeline, a 300 MW coal-fired electric power generation plant, power transmission lines and a large tailings management facility.

6. Over 150 permits have been identified as being required to develop the Mina de Cobre Panamá project.

7. The total estimated cost to design and build the Mina de Cobre Panamá project (150,000 t/d) is US$4.32 billion. Expansion of the concentrator in Year 10 to a nominal capacity of 225,000 t/d and other sustaining capital over the life of the project is estimated at an additional US$1.72 billion.

8. Operating costs, including mining, processing, site services and G&A, are estimated at approximately US$7.23/t of milled ore.




9. The un-levered, after-tax internal rate of return for the project is estimated at 12.4%. A levered case, assuming 50% debt financing totaling US$2.16 billion, yields an after-tax 15.1% internal rate of return. The payback period is projected at 5.9 years.

The FEED Study has successfully expanded the mineral resource and mineral reserve base of the project to support the proposed development plan. The Mina de Cobre Panamá project is technically and economically feasible.




20.0 RECOMMENDATIONS

The following actions are recommended for the Mina de Cobre Panamá project:

1. The project should proceed to the next phase of basic engineering for subsequent development in anticipation of receiving the necessary permits.

2. Inmet Mining and its wholly-owned subsidiary Minera Panamá S.A. should complete the currently underway Estudio de Impacto Ambiental Category III (ESIA, or Environmental and Social Impact Assessment) and proceed with acquiring the necessary permits for project construction and operation.

3. Additional in-fill drilling should be performed in Colina mining phases 1 and 2 and Valle Grande mining phase 1 to upgrade the classification of mineral reserves for material that will likely be mined during Years 6-10 of the project. The work should be completed during late preproduction through the first 2-3 years of operation.

4. Conduct additional drilling to potentially convert inferred mineral resources within the proposed pit limits to mineral reserves. This could extend the life of the project and reduce the stripping ratio. The drilling would be in peripheral areas of each deposit that would be developed in later years; consequently, its completion is not urgent. Botija should be targeted first as its development precedes the other deposits. It should be remembered that inferred mineral resources are speculative and uncertain in nature and cannot be relied upon. There is no assurance that inferred mineral resources will be upgraded to mineral reserves.

5. There are a number of exploration targets such as Brazo, Botija Abajo and extensions of the Colina and Valle Grande deposits that have the potential to expand the life of the project and should be further tested by exploration drilling. Brazo and the northeast extension of Colina could impact mining plans during the first 10-15 years of operation. The Brazo area should be drilled in late preproduction through the first 2-3 years of operation as indications of higher grades may affect early mine development.

6. Additional exploration, including a ZTEM airborne survey, soil geochemical surveys and diamond drilling, is recommended to test the western extent of the Cobre Mineral trend.

7. Engineering studies should be conducted to optimize the type and location of the Colina crusher and related conveyor facilities, which must be placed into operation by the start of Year 6. This includes foundation and other geotechnical testing, exploration/condemnation drilling in proposed sites, and economic trade-off analyses. Consideration should be given to a semi-mobile facility that could be moved periodically to minimize ore haulage distances from the Colina and Valle Grande deposits and, therefore, potentially reduce mine operating costs. These studies should




be performed during the preproduction development period through the first two years of operation.




21.0 REFERENCES

AMEC, 2010. Mina de Cobre Panamá Project FEED Study Report; AMEC Report AMEC, 2007. Petaquilla Project, Panamá NI 43-101 Technical Report; AMEC Report, 174 p. Awmack, H.J., 1992. Qualifying Report on the Petaquilla Project; Adrian Resources Limited, Internal Report, 25 p. Awmack, H.J., 1992. Summary Report on the Petaquilla Project; Inmet Mining Corportation, Minera Panamá S.A. Awmack, H.J., and G.F. McArthur, 1993. 1993 Interim Report on the Petaquilla Project; Adrian Resources Limited, Internal Report. CIM, 2005. CIM Definition Standards - For Mineral Resources and Mineral Reserves; 11 December 2005, Prepared by the CIM Standing Committee on Reserve Definitions. Cruden, A., 1998. On the emplacement of tabular granites, Journal of the Geological Society, London. Vol. 155, p. 853-862. Davis, B.M., 1997. Some Methods of Producing Interval Estimates for Global and Local Resources, SME Preprint 97-5, 4p. de Boer J.Z., M.S. Drummond, M.J. Bordelon, M.J. Defant, H. Bellon, and R.C. Maury, 1995. Cenozoic Magmatic Phases of the Costa Rican Island Arc (Cordillera de Talamanca); in Geologic and Tectonic Development of the Caribbean Plate Boundary in Southern South America, Mann, P, ed.; Geological Society of America, Special Paper 295, p. 35-56.

Escalante, A., 2009. Overview of the Geology of the Botija porphyry Cu-Au-Mo deposit, Panamá. Minera Panamá S.A.. Internal Report.

Francois-Bongarcon, D., 1995. Petaquilla – Check Assays; 13 October 1995, Mineral Resources Development Inc. Memorandum to J. Nilsson, B. McEwan, B. Putnam, and T. Eggleston.

Gustafson, L.B. and Hunt, J.P., 1975. The porphyry copper deposit at El Salvador, Chile. Economic Geology, v. 70, p. 857-912. Journel, A.G. and Ch.J. Huijbregts, 1978. Mining Geostatistics; Academic Press, 600 p.

Kesler, S.E., J.F. Sutter, J.J. Issigonis, L.M. Jones, and R.L. Walker, 1977. Evolution of Porphyry Copper Mineralization in an Oceanic Island Arc: Panamá; Economic Geology, v. 72, p. 1142-1153.




McArthur, G.F., S. Harris, S. Kenwood, and D. Laudrum, 1995. 1994 Summary Report on the Petaquilla Project; Adrian Resources Limited, Internal Report.

Mann, P. 1995. Preface: Geologic and Tectonic Development of the Caribbean Plate Boundary in Southern South America. Geological Society of America, Special Paper 295, p. xi-xxxii.

Panamá Mineral Resources Development Company, 1977. Preliminary Feasibility Report of the Petaquilla Project; PMRD Internal Report, 42 p.

Sides, E.J., 1994. Effect of Barren Dikes on Ore Potential at Cerro Colorado Porphyry Copper Deposit, Panamá; Transactions of the Institute of Mining and Metallurgy, v. 93, p. B39-B47.

Simons 1996. Petaquilla Project Feasibility Study. Issued to Teck Cominco, November 1996. Simons 1998. Petaquilla Project Feasbility Study. Issued to Teck Cominco, January 1998. Speidel, F. and Faure, S., 1996. Surface Geology of Botija and Petaquilla deposits. Panamá, Internal Report, 36 p. Speidel, F., Faure, S., Smith, M.T., McArthur, G.F., 2001. Exploration and discovery at the Petaquilla Copper-Gold Concession, Panamá. Soc. Econ. Geo. Spec. Pub. 8, p. 349-362. Thompson, A.J.B., 1996. Petaquilla project, Panamá: Petrographic descriptions and whole rock geochemistry, Petrascience Consultants Inc., unpublished report for Inmet Mining Corporation. Twiss, R.J. and Moores, E.M., 1992. Structural Geology. W.H. Freeman and Company, New York, p. 1-532. United Nations Development Program, 1969. Porphyry Copper Mineralization at Cerro Petaquilla, Province of Colón, Panamá; United Nations Development Program Report, 87 p.

Žák, J.; Holub, F.V.; Kachlíl, V., 2006. Magmatic stoping as an important emplacement mechanism of Variscan plutons: evidence from roof pendants in the Central Bohemian Plutonic Complex (Bohemian Massif). International Journall of Earth Sciences (Geol Rundsch). 95, p. 771-789.

Mesoamerican Biological Corridor, A platform for sustainable development, Technical Series 02, Project for the Consolidation of the Mesoamerican Biological Corridor 2002, http://www.ccad.ws/PCCBM/docs/platform.pdf

Millennium Development Goals: Panamá http://www.indexmundi.com/panama/millennium-development-goals.html




22.0 DATE AND SIGNATURES

The effective date of this report is March 31, 2010.

The execution date (date of signing) of this report is May 3, 2010.




CERTIFICATE of QUALIFIED PERSON

I, William L. Rose, P.E., do hereby certify that:

1. I am currently employed as the Principal Mining Engineer by:

WLR Consulting, Inc. 9386 West Iowa Avenue Lakewood, Colorado 80232-6441 U.S.A.

2. I graduated with a Bachelor of Science degree in Mining Engineering from the Colorado School

of Mines in 1977. 3. I am a:

Registered Professional Engineer in the State of Colorado (No. 19296) Registered Professional Engineer in the State of Arizona (No. 15055) Registered Member of the Society for Mining, Metallurgy and Exploration, Inc.

(no. 2762350RM) 4. I have practiced my profession continuously for 32 years since my graduation from college.

My experience includes underground and open pit projects in coal, industrial minerals, base and precious metals, and includes performing mineral resource and reserve estimations, open pit mine design and planning, production scheduling, estimations of mining equipment and manpower requirements, estimations of mine capital and operating costs for scoping-level, prefeasibility and feasibility studies for projects in North America, Central and South America, the Philippines, Australia, Africa and Europe.

5. I have read the definition of “Qualified Person” set out in National Instrument 43-101 (“NI 43-

101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “Qualified Person” for the purposes of NI 43-101.

6. I am responsible for the overall preparation of the technical report titled “Mina de Cobre

Panamá Project, Panamá NI 43-101 Technical Report” effective date March 31, 2010 (the “Technical Report”), prepared for Inmet Mining Corporation. and, in particular, am responsible for the preparation of those portions of the Technical Report not specifically prepared by other qualified persons.

7. I have not had prior involvement with the property that is the subject of the Technical Report. I

have visited the subject property on April 15-16, 2009. 8. As of the date of this certificate, to the best of my knowledge, information and belief, the

Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.




9. I am independent of the issuer applying all of the tests in Section 1.4 of National Instrument 43-101.

10. I have read National Instrument 43-101 and Form 43-101F1, and the Technical Report has

been prepared in compliance with that instrument and form. 11. I consent to the filing of the Technical Report with any stock exchange and other regulatory

authority and any publication by them for regulatory purposes, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report.

Dated this 3rd day of May, 2010.

Signed and Sealed

Signature of Qualified Person

William L. Rose

Print Name of Qualified Person





I, Colin M Burge, P.Geo., do hereby certify that:

1. I am currently employed as a Senior Geologist by:

Inmet Mining Corporation 330 Bay Street, Suite 1000 Toronto, Ontario M5H 2S8

2. I graduated with a Bachelor of Science degree in Earth Science from University of Waterloo in

1981. 3. I am a member of the Association of Professional Engineers and Geoscientists of British

Columbia, registration number 20274 4. I have worked as a geologist for 25 years after graduation from university. 5. I have read the definition of “Qualified Person” set out in National Instrument 43-101 (“NI 43-

101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “Qualified Person” for the purposes of NI 43-101 but do not fulfill the requirements of an “independent qualified person”.

6. I have assisted in the preparation of the technical report titled “Mina de Cobre Panamá,

Project, Panamá NI 43-101 Technical Report” effective date March 31 2010 (the “Technical Report”), prepared for Inmet Mining Corporation. and, in particular, I prepared Sections 7 through 13 and parts of 19 and 20 of the Technical Report.

7. I have had prior involvement with the property that is the subject of the Technical Report. The

nature of my prior involvement is that I personally supervised technical [field] work on the property from July 1, 2007 to the present.

8. As of the date of this certificate, to the best of my knowledge, information and belief, the




authority and any publication by them for regulatory purposes, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report.





Signed and Sealed Signature of Qualified Person

Colin Michael Burge






I, Gary S. Wells, P.Geo., do hereby certify that:

1. I am currently employed as a Senior Geologist by:

Inmet Mining Corporation 330 Bay Street, Suite 1000 Toronto, Ontario M5H 2S8

2. I graduated with a Bachelor of Science degree in Geology and Chemistry from Carleton

University in 1975 and a Ph.D. in Geology from Queen’s University in 1980. 3. I am a member of the Association of Professional Engineers and Geoscientists of British

Columbia, registration number 19849 4. I have worked as a geologist for 30 years since my graduation from university. 5. I have read the definition of “Qualified Person” set out in National Instrument 43-101 (“NI

43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “Qualified Person” for the purposes of NI 43-101 but do not fulfill the requirements of an “independent qualified person”.

6. I have assisted in the preparation of the technical report titled “Mina de Cobre Panamá

Project, Panamá NI 43-101 Technical Report” effective date March 31, 2010 (the “Technical Report”), prepared for Inmet Mining Corporation and, in particular, I prepared sections 4 through 13, section 15, parts of sections 19 and 20, and section 21 of the Technical Report.

7. I have had prior involvement with the property that is the subject of the Technical Report.

The nature of my prior involvement was to maintain and update the geological and digital drilling database for the project since October 2008 and to construct the geological models used in the resource modeling.

8. I have visited the property on several occasions since 2008, most recently from November

14th to 17th, 2009.




been prepared in compliance with that instrument and form.




11. I consent to the filing of the Technical Report with any stock exchange and other

regulatory authority and any publication by them for regulatory purposes, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report.


Signed and Sealed

Signature of Qualified Person

Gary S. Wells






I, Bruce Davis, FAusIMM, do hereby certify that:

1. I am an independent consultant of BD Resource Consulting, Inc., located at 4253 Cheyenne Drive, Larkspur, CO, U.S.A., 80118, incorporated January 18, 2008.

2. I graduated with a Doctor of Philosophy degree from the University of Wyoming in 1978. 3. I am a fellow of the Australasian Institute of Mining and Metallurgy, Registration Number 2111185. 4. I have practiced my profession continuously for 32 years and have been involved in geostatistical

studies, mineral resource and reserve estimations and feasibility studies on numerous underground and open pit base metal and gold deposits in Canada, the United States, Central and South America, Europe, Asia, Africa and Australia.

5. I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and

certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.

6. I am responsible for the QA/QC evaluations in Section 14 and for the geostatistical analyses

described in Section 17 of the Technical Report titled “Mina de Cobre Panamá Project, Panamá NI 43-101 Technical Report” with effective date March 31, 2010, (the “Technical Report”). I personally visited the site from April 14 to April 18, 2009.

7. During the period between 1996 and 1998 I was involved in the project in the position of

Geostatistician, with H. A. Simons during delineation drilling and subsequent feasibility studies conducted by H. A. Simons for Adrian Resources and Teck Corporation.

8. As of the date of this certificate, to the best of my knowledge, information and belief, the Technical

Report contains all scientific and technical information that is required to make the Technical Report not misleading.

9. I am independent of the issuer applying all of the tests in Section 1.4 of National Instrument 43-101. 10. I have read National Instrument 43-101 and Form 43-101F1, and the Technical Report has been

prepared in compliance with that instrument and form. 11. I consent to the filing of the Technical Report with any stock exchange and other regulatory

authority and any publication by them, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report

Dated this 3rd Day of May , 2010. _ Bruce M. Davis, FAusIMM





I, Robert Sim, P.Geo, do hereby certify that:

1. I am an independent consultant of SIM Geological Inc., located at 6810 Cedarbrook Place, Delta, BC, Canada, V4E 3C5, incorporated December 20, 2005 (BC 0743802).

2. I graduated from Lakehead University with an Honours Bachelor of Science (Geology) in 1984. 3. I am a member of the Association of Professional Engineers and Geoscientists of British

Columbia, License Number 24076. 4. I have practiced my profession continuously for 26 years and have been involved in mineral

exploration, mine site geology and operations, mineral resource and reserve estimations and feasibility studies on numerous underground and open pit base metal and gold deposits in Canada, the United States, Central and South America, Europe, Asia, Africa and Australia.

5. I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-

101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.

6. I am responsible for the preparation of the mineral resources estimates described in Section

17.1 of the Technical Report titled “Mina de Cobre Panamá Project, Panamá NI 43-101 Technical Report“ with effective date March 31, 2010, (the “Technical Report”). I personally visited the site from April 14 to April 18, 2009.

7. During the period between 1996 and 1998 I was involved in the project in the position of Senior

Geologist, Resource Evaluations with Inmet Mining Corporation. I was one of Inmet’s technical representatives during delineation drilling and subsequent feasibility studies conducted by Adrian Resources and Teck Corporation.


Technical Report contains all scientific and technical information that is required to make the Technical Report not misleading.

9. I am independent of the issuer applying all of the tests in Section 1.4 of National Instrument 43-

101. 10. I have read National Instrument 43-101 and Form 43-101F1, and the Technical Report has


authority and any publication by them, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report




Dated this 3rd Day of May , 2010.

_ Robert Sim, P. Geo.




CERTIFICATE OF QUALIFIED PERSON

Alexandra J. Kozak, P.Eng. AMEC Americas Limited

111 Dunsmuir Street, Suite 400 Vancouver, BC

Tel: (604) 664-4578 Fax: (604) 664-3057

[email protected]

I, Alexandra J. Kozak, P.Eng., am employed as Manager, Process Engineering with AMEC Americas Limited.

This certificate applies to the Technical Report entitled “Inmet Mining Corporation, Mina de Cobre Panamá Project, Panamá, NI 43-101 Technical Report” (the Technical Report) dated March 31, 2010.

I am a member of the Association of Professional Engineers and Geoscientists of British Columbia (APEGBC). I graduated from the University of Alberta with a Bachelor of Science degree in Mineral Process Engineering in 1985.

I have practiced my profession continuously since 1985 and have been involved in operations in Canada and Guyana and preparation of scoping, pre-feasibility, and feasibility level studies for gold, base metals and diamond properties in Canada, United States, Peru, Mexico, Mongolia, Ghana, and New Guinea. I am currently a Consulting Engineer and have been so since September 1996.

As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43–101 Standards of Disclosure for Mineral Projects (NI 43–101).

I have not visited the Mina de Cobre Panamá Project.

I am responsible for Section 16, 23.2, 23.3, 23.5.2, 23.5.3 and those portions of the Summary, Interpretation and Conclusions, and Recommendations that pertain to those sections, of the Technical Report.

I am independent of Inmet Mining Corporation as independence is described by Section 1.4 of NI 43-101.

I have been involved with the Mina de Cobre Panamá Project since March 2010, as part of the preparation of this Technical Report.

I have read NI 43–101 and this Technical Report has been prepared in compliance with that Instrument.




As of the date of this certificate, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Signed and Sealed

Alexandra J. Kozak, P.Eng.

Dated: 3 May 2010

AMEC Americas Limited 111 Dunsmuir Street, Suite 400 Vancouver, B.C. V6B 5W3 Tel (604) 664-3030 Fax (604) 664-3057 www.amec.com




23.0 ADDITIONAL REQUIREMENTS FOR TECHNICAL REPORTS ON DEVELOPMENT PROPERTIES AND PRODUCTION PROPERTIES

23.1 Mining Operations

A mine production schedule was developed from quarterly, annual and multi-year open pit sequence plans using a variable NSR cutoff grade strategy to increase early revenues to improve overall project returns. Cutoffs will be elevated significantly during the first seven years of operations and then gradually decline to internal cutoffs around Year 14. A third grinding circuit will be added to the concentrator and will commence operation in Year 10, increasing the base ore processing rate from 150,000 t/d to 225,000 t/d. Owner’s preproduction stripping would commence about 15 months before plant start-up.

Table 23-1 summarizes the resulting mine production and material handling schedule. Mine operations will be scheduled for two 12-hour shifts per day, 365 days per year. The concentrator will operate an estimated 29.5 years, including the processing of about 173 Mt of stockpiled ore during Years 28 to 30. During preproduction stripping and pit operations, approximately 65 Mt of saprock and 119 Mt of low-grade ore, totalling 184 Mt, will be stockpiled in selected areas within the Botija West area. The ROM stockpile is estimated to contain about 1.5 Mt. Stockpile recovery losses are estimated at 7% to account for inaccessibility along steep terrain and material losses in the contact zone with felled timber.

Table 23-1: Mine Production Schedule

Time Period

Ore to ROM Stockpile or Mill

(kt)

To Saprock Ore Stockpile

(kt)

To Low-Grade Ore Stockpile

(kt)

Waste Rock & Saprolite

(kt) Total Material

(kt) Strip Ratio Contractor

(kt) Owner

(kt)

Prior to M-15 166 1,279 402 55,567 57,414 344.87 57,414 -

PP M-15 to M0 1,306 5,745 4,933 48,902 60,886 45.62 10,021 50,865

Y1 42,445 3,548 16,857 64,674 127,524 2.00 6,729 120,795

Y2 58,187 3,127 20,900 39,745 121,959 1.10 - 121,959

Y3 57,221 - 22,892 25,493 105,606 0.85 2,096 103,510

Y4 57,106 2,504 18,880 34,849 113,339 0.98 10,435 102,904

Y5 55,906 11,425 10,913 31,349 109,593 0.96 6,368 103,225

Y6 57,104 4,610 5,353 40,218 107,285 0.88 8,505 98,780

Y7 56,811 10,782 5,738 52,040 125,371 1.21 12,571 112,800

Y8 56,676 3,725 3,607 60,599 124,607 1.20 7,316 117,291

Y9 54,800 4,220 4,127 57,190 120,337 1.20 - 120,337

Y10 80,187 239 4,365 53,717 138,508 0.73 - 138,508

Y11-Y15 413,003 3,066 - 294,492 710,561 0.72 20,355 690,206

Y16-Y20 420,392 8,484 - 269,593 698,469 0.66 44,664 653,805

Y21-Y25 369,396 2,344 - 135,213 506,953 0.37 - 506,953

Y26-Y30 * 363,418 - - 24,392 387,810 0.07 - 387,810

Total 2,144,124 65,098 118,967 1,288,033 3,616,222 0.69 186,474 3,429,748

* Includes 172,549 kt of ore stockpile reclamation in Years 28 to 30. The 1,472 kt of ROM ore stockpiled during preproduction is therefore double counted in the total line. Mineral reserves are estimated at 2,142,652 kt.




The contractor’s construction materials excavation, saprolite removal, and limited waste rock stripping are estimated at about 186 Mt over the life of the mine. Of this total, about 149 Mt is saprolite. Under the schedule in Table 23-1, the Owner will handle more than 3.43 billion tonnes (Gt) during the 31 years of preproduction stripping and mine operations.

The contractor will excavate about 37 Mt from a construction materials quarry before Month -15. Of this total, about 5 Mt is saprolite, 1.8 Mt is ore-grade material that will be stockpiled, and about 30 Mt is waste rock – half of which is projected to be low-sulphur material suitable for construction of other project facilities.

The following Owner’s primary equipment will be required for the peak mining rates during Years 11 to 20:

(a) 7 blasthole drills ........................................................ 311 mm diameter, 60 t bit loading (b) 4 electric shovels ...................................................... 55 m3 (c) 2 front-end loaders ................................................... 38 m3 (d) 36 off-highway haul truck ......................................... 360 t payload (e) 2 crawler dozers ....................................................... 635 kW (D11-class) (f) 8 crawler dozers ....................................................... 435 kW (D10-class) (g) 4 rubber-tired dozers ................................................ 370 kW (834H-class) (h) 2 motor graders ........................................................ 400 kW (24M-class) (i) 3 motor graders ........................................................ 220 kW (16M-class) (j) 3 water trucks ........................................................... 80,000 litre (96 t)

Mine workforce levels will vary between about 317 and 563 people during the operating years, depending on production rates and haulage distances. This includes both salaried and hourly workers, expatriates, and nationals. Four rotating crews will provide continuous operator and maintenance coverage in the mine.

Table 23-2 presents the projected mill feed schedule for the FEED Study mine plan. RCu, RMo and RAu refer to recoverable grades of Cu, Mo and Au, respectively. Ag Rec refers to the anticipated recovery for Ag expressed in percent. HrPKt refers to milling hours per kilo-tonne of ore at a nominal grinding rate of 150,000 t/d.




Table 23-2: Mill Feed Schedule (all ore types)

Proven + Probable Mineral Reserves Milled Year kt NSR $/t Cu% Mo% Au g/t Ag g/t HrPKt RCu % RMo % RAu g/t Ag Rec % Y1

Y1 Q1 5,775 15,04 0.41 0.010 0.08 1.31 0.160 0.36 0.006 0.05 47.3 Y1 Q2 10,580 17.47 0.47 0.008 0.11 1.35 0.152 0.42 0.005 0.07 47.3 Y1 Q3 12,739 18,99 0.52 0.009 0.11 1.30 0.162 0.46 0.006 0.07 47.3 Y1 Q4 13,351 19.69 0.53 0.010 0.11 1.53 0.159 0.48 0.006 0.07 47.3

Total Y1 42,445 18.29 0.50 0.009 0.11 1.39 0.158 0.44 0.006 0.06 47.3 Y2

Y2 Q1 14,857 20.40 0.55 0.010 0.12 1.43 0.150 0.49 0.007 0.07 47.3 Y2 Q2 14,674 20.75 0.56 0.009 0.12 1.42 0.152 0.51 0.006 0.07 47.3 Y2 Q3 14,056 19.66 0.53 0.010 0.11 1.31 0.160 0.48 0.006 0.07 47.3 Y2 Q4 14,600 20.62 0.56 0.010 0.11 1.51 0.151 0.51 0.006 0.06 47.3

Total Y2 58,187 20.37 0.55 0.010 0.11 1.42 0.153 0.50 0.006 0.07 47.3 3 57,221 20.73 0.56 0.011 0.11 1.41 0.152 0.50 0.007 0.07 47.3 4 57,106 20.67 0.56 0.010 0.11 1.46 0.150 0.51 0.006 0.07 47.3 5 55,906 22.00 0.59 0.011 0.11 1.61 0.154 0.54 0.007 0.07 47.3 6 57,104 21.88 0.59 0.011 0.11 1.69 0.155 0.54 0.007 0.07 47.3 7 56,811 20.61 0.56 0.008 0.14 1.74 0.154 0.50 0.005 0.09 47.3 8 56,676 19.23 0.53 0.009 0.09 1.73 0.153 0.47 0.006 0.06 47.3 9 54,800 18.74 0.52 0.010 0.09 1.67 0.157 0.46 0.006 0.05 47.3 10 80,187 15.47 0.43 0.008 0.08 1.48 0.155 0.38 0.005 0.04 47.3 11 82,601 14.22 0.41 0.007 0.06 1.37 0.155 0.35 0.004 0.03 47.3 12 82,601 14.22 0.41 0.007 0.06 1.37 0.155 0.35 0.004 0.03 47.3 13 82,601 14.22 0.41 0.007 0.06 1.37 0.155 0.35 0.004 0.03 47.3 14 82,600 14.22 0.41 0.007 0.06 1.37 0.155 0.35 0.004 0.03 47.3 15 82,600 14.22 0.41 0.007 0.06 1.37 0.155 0.35 0.004 0.03 47.3 16 84,078 14.22 0.40 0.008 0.08 1.39 0.155 0.35 0.005 0.04 47.3 17 84,078 14.22 0.40 0.008 0.08 1.39 0.155 0.35 0.005 0.04 47.3 18 84,078 14.22 0.40 0.008 0.08 1.39 0.155 0.35 0.005 0.04 47.3 19 84,079 14.22 0.40 0.008 0.08 1.39 0.155 0.35 0.005 0.04 47.3 20 84,079 14.22 0.40 0.008 0.08 1.39 0.155 0.35 0.005 0.04 47.3 21 73,879 11.97 0.35 0.006 0.05 1.50 0.173 0.30 0.004 0.03 47.3 22 73,879 11.97 0.35 0.006 0.05 1.50 0.173 0.30 0.004 0.03 47.3 23 73,879 11.97 0.35 0.006 0.05 1.50 0.173 0.30 0.004 0.03 47.3 24 73,879 11.97 0.35 0.006 0.05 1.50 0.173 0.30 0.004 0.03 47.3 25 73,880 11.97 0.35 0.006 0.05 1.50 0.173 0.30 0.004 0.03 47.3 26 81,120 10.12 0.32 0.006 0.05 1.29 0.162 0.24 0.003 0.02 37.5 27 81,120 10.12 0.32 0.006 0.05 1.29 0.162 0.24 0.003 0.02 37.5 28 81,120 10.12 0.32 0.006 0.05 1.29 0.162 0.24 0.003 0.02 37.5 29 81,120 10.12 0.32 0.006 0.05 1.29 0.162 0.24 0.003 0.02 37.5 30 38,938 10.12 0.32 0.006 0.05 1.29 0.162 0.24 0.003 0.02 37.5 Total 2,142,652 14.60 0.41 0.008 0.07 1.43 0.159 0.36 0.005 0.04 45.8

Note: Estimated from Botija deposit model of January 2010 and Colina-VG deposit model of November 2009.

23.2 Ore Processing

Ore from the Botija, Colina, and Valle Grande pits will be treated in a large concentrator using current technology to produce a copper concentrate and a molybdenum concentrate for sale on the world market. The concentrator will initially treat a nominal 150,000 t/d of ore supplied from the Botija pit; later, ore will be received from the Colina and Valle Grande pits. From Year 10,




the concentrator ore throughput will be increased by 50%, to a nominal 225,000 t/d, to maintain production of concentrate despite a falling head grade. Crushing, grinding, bulk rougher flotation, water, and air systems will increase in capacity by 50% to accomplish the increase in ore treatment rate; all other systems will remain at the same size.

The process plant is designed to process ore at a head grade of 0.7% Cu and 0.013% Mo. These levels are higher than the highest sustained head grades of 0.58% Cu and 0.011% Mo expected to be mined in Year 5, but the design provides the flexibility to accommodate a wide range of head grades over the project life. The plant design also allows for 15% day-to-day fluctuations in throughput. The process includes the following facilities:

crushing and grinding to liberate minerals from the ore froth flotation to separate most of the copper and molybdenum minerals from minerals

of no commercial worth differential flotation to separate the copper and molybdenum minerals from each other facilities to store tailings and provide reclaim water for the process facilities to remove water from the products and to ship concentrates to market.

Run-of-mine (ROM) ore will be delivered by haul truck to the dump pockets of two primary gyratory crushers installed in a single in-ground concrete structure close to the rim of the Botija pit. A 400,000 tonne ROM stockpile will be located close to the crushers to provide a 2½-day supply of ore for times when weather conditions preclude hauling ore out of the pit. The ROM stockpile will be operated on a first in, first out basis to prevent the accumulation of aged ore.

Separate feeders and take-away conveyors will move the ore from each crusher to a series of conveyors which will discharge the ore onto a conical coarse ore stockpile at the concentrator. There will be provision at the transfer point between the two overland conveyors to accept mill feed from future crushed ore sources. The coarse ore stockpile will hold a 2½-day supply for the mill, 15 hours of which will be available to the reclaim feeders without the assistance of a bulldozer.

Two trains of feeders and conveyors will draw ore from below the coarse ore stockpile and feed two parallel wet-grinding lines, each consisting of a semi-autogenous grinding (SAG) mill and two ball mills, all equipped with gearless drives. The SAG mill circuits will be closed by trommel screens followed by washing screens; conveyors will deliver screen oversize to pebble crushers. The pebble crushing circuits will include pebble bins, cone crushers, and a bypass arrangement. Crushed pebbles will return to the SAG mills via the feed conveyors. From Year 10 of operation, another coarse-ore stockpile and grinding line will be added to increase the ore treatment rate.

Discharge from each SAG mill will be evenly split between two ball-mill circuits. The four ball-mill circuits will be closed by hydrocyclones. Ground slurry will be directed to a flotation circuit where a bulk sulphide concentrate, containing copper, molybdenum, and gold values, will be collected and concentrated in a rougher followed by three stages of cleaner flotation. The roughers and first cleaners will be tank cells, while the second and third cleaners will be column




cells. Before cleaning, rougher concentrate will be reground in vertical stirred mills. From Year 10, a 50% increase in rougher capacity will be required to accommodate the increase in throughout, but the amount of copper will be the same; therefore, no change to the existing downstream regrind and cleaning capacity will be needed.

When the molybdenum head grade warrants operating the molybdenum plant, the bulk concentrate will be thickened in a conventional thickener (with no flocculant) and pumped to a differential flotation plant, where copper minerals will be depressed and molybdenite floated into a molybdenum concentrate. The molybdenum concentrate will be filtered, dried, and packaged in tote bags for shipment to offshore roasters. Tailings from the molybdenum flotation circuit will constitute the copper concentrate, which will be pumped approximately 30 km to a filter plant at a port on the Caribbean coast. If the molybdenum head grade is very low, the molybdenum separation plant will be bypassed.

A simplified flow diagram is provided in Figure 23-1.




Figure 23-1: Process Flow Diagram

Primary Crusher

Open Pit Mine

SAG Mills (2)

Screens

Coarse Ore

Stockpile

Rougher Flotation (4 banks of 7)

3rd Cleaner Flotation (4)

2nd Cleaner Flotation (6)

1st Cleaner Flotation (2 banks of 8)

1st Mo Clnr

Flotation

Cu Conc.Thickener

Cu ConcStorage Tank

(2)

MainlinePumps (2x5)

Copper Filters (4)

Cu Conc. Loadout

Pipelines (32 km)

Tailings to Impoundment

Regrind Vertimills

(4)

Cu Conc.Storage Tank (2)

Ball Mills (4)

Cyclone Cluster (4)

CycloneCluster (2) Bulk Conc.

Thickener

Bulk Concentrate

Storage Tanks (2)

Mo Conditioner

o/f

Potentially Acid-Generating Tailings to Impoundment

To Process Water

Shiploader

Mo Rougher Flotation Mo Scav.

Flotation

2nd Mo Clnr Flotation

3rd Mo Clnr Flotation



Mo Conc.PackingMo Dryer

BinsMoFilter

Regrind Vertimill

Pebble Crushers (2)

Mo Concentrate Thickener

Thickener/Clarifier

Overflow

Port

MineSite

The locations of the process plant and ancillary facilities at the mine/plant site are shown in Figure 23-2. The facilities are centrally located with respect to the three open pits and associated WRSFs. The available area is rather limited, but the layout has been designed for efficient material flow and personnel access to the buildings.

The facilities will be constructed along ridgelines or on hill tops to maximize the use of gravity for milling, flotation, and tailings discharge, to minimize cut-and-fill quantities, and to place the structures on competent ground.




Figure 23-2: Layout of Process Plant Area

Primary Crushers

ROM Stockpile

Service Vehicle Shop

Coarse Ore Stockpile

Truck Shop

Main Substation

Flotation

Reagents

Bulk Thickener

Copper Thickener

Site of Future Stockpile

Site of Future 3rd Line

Plant Maintenance Shop

Process Water Pond

Process Water Tanks

Laboratory

23.2.1 Tailings Disposal

Rougher tailings, found in testwork to be non acid-generating, will be piped by gravity to a cyclone house at the south end of the TMF. Coarse material in the tails will be used to construct sand dams to contain the tailings; about half of the coarse tailings will need to be cycloned to provide enough material. Cyclone underflow (sand) will be spigotted onto the downstream faces of the dam, and cyclone underflow and uncycloned tailings will be distributed around the upstream faces of the dams and onto the beaches of the TMF. Cleaner tails are




presumed to be potentially acid generating and will be directed by gravity in a separate line for subaqueous discharge in the TMF.

Direct precipitation into the TMF, together with contact water directed to the facility from the WRSFs and other disturbed land in the area, will result in a net excess of water in the TMF. Some of this water will be recycled back to the process plant by barge-mounted pumps. Excess water will be discharged at a controlled rate to the natural watercourses downstream of the impoundment, via a polishing pond, and will meet all applicable emission limit values and receiving water quality criteria.

23.3 Metallurgical Recoveries and Concentrator Production


In 1997, an extensive program of metallurgical testing was designed to confirm earlier work on the metallurgical response of the Botija and Colina deposits. Most of the work was done at Lakefield Research Ltd., Lakefield, Ontario. Grinding, flotation, dewatering, and mineralogical work were performed as part of this program. In addition to the Lakefield work, locked-cycle flotation testwork and modal analysis were performed at G&T Metallurgical Services Ltd., Kamloops, B.C. (G&T) to assist in defining grind requirements for both rougher and cleaner flotation. Copper-molybdenum separation by differential flotation was conducted by International Metallurgical and Environmental, Kelowna, B.C. (IME). The metallurgical work done for the present study has built upon the 1997/1998 study with some knowledge of, but no reliance on, work performed before that time.

The testwork before 2007 was based on large composite samples, and the results, particularly for flotation testing, could not be used for interpreting the variability of response for material within the deposits. Consequently, a large sampling program was undertaken in 2008/2009 to bolster the knowledge from previous work and provide the missing insight into the variability of response. A total of 16 special holes for metallurgical grinding and flotation tests were drilled in the Botija, Valle Grande, and Colina orebodies. Sample preparation, flotation testing, and testing of flotation products were done primarily at G&T. Grinding work was conducted at SGS Mineral Services, Lakefield, Ontario, and at Philips Enterprises LLC, Golden, Colorado.The life-of mine production and recovery data are shown in Table 23-3. Recovery formulas for each metal are described in more detail in Section 17.2.2.







t/a 71,968,053 73,281,400 71,421,733

t/d 197,173 200,771 195,676

Head Grade % Cu 0.46 0.34 0.41

% Mo 0.008 0.006 0.008

g/t Au 0.08 0.05 0.07

oz/t Ag 1.46 1.39 1.43

Recovery % Cu 88.4 79.5 85.9

% Mo 61.9 53.1 59.0

% Au 57.2 44.6 54.3

% Ag 47.3 42.8 45.8


% Cu 28 28 28


t/a 7,090 4,741 6,188

% Mo 52 52 52

23.4 Infrastructure

Various project support facilities will be provided at the mine/plant site and the port site. The mine site facilities are divided into two areas: the mine/plant site, which includes buildings and structures for repair and maintenance of mine and plant equipment, and the eastern infrastructure area, which includes facilities for personnel accommodations, administration, and security. The port site includes facilities for concentrate storage and load-out to ocean-going vessels, coal receiving facilities, a barge berth, and inbound/export freight handling and storage facilities. The right-of-way for the access corridor between the two sites will be occupied by a new road, pipelines, and power transmission lines.

Proposed layouts of the mine/plant site, eastern infrastructure area, and port site are shown in Figure 23-3, 23-4, and 23-5.




Figure 23-3: Layout of Mine/Plant Site




Figure 23-4: Layout of Eastern Infrastructure Area




Figure 23-5: General Arrangement of Port Site Facilities




23.4.1 Mine/Plant Site

Support infrastructure for the plant and mine operations includes a truckshop, surface vehicle shop, mill maintenance shop, fuel storage depot, substation, container storage and laydown, facilities for storing and preparing blasting agents, process water pond, process and fire water tanks, and site roads.

The mine truckshop will be the main services complex on site, containing maintenance facilities for the mine mobile equipment fleet, warehouse space, a machine shop, offices, and the mine dispatch centre.

A mine service vehicle shop will be provided for all light- and medium-duty vehicles such as non-mining mobile equipment, road transport vehicles, personnel buses, and pickup trucks.

Facilities for a licensed blasting contractor will be provided on the opposite side of a ridge south of the plant site area. The foundations will be cut into the hillside to provide a natural barrier to protect the mine/plant facilities as well as the nearby Molejón mine property to the southeast.

23.4.2 Eastern Infrastructure Area

The eastern infrastructure area, about 4 km to the east of the plant site, will be used for the mine site construction camp, the permanent camp, the administration building, the construction power plant, and the 230 kV switchyard.

The construction camp will be sized for 3,500 workers and the permanent camp for 1,200 workers, including those employed at the port site during operations. A separate 900-person construction camp will be provided at the port site.

The main administration building will include a reception area, a small training room, and other standard office features. An elevator will be provided for full accessibility. A separate training facility adjacent to the permanent camp will be equipped to provide all training functions for the mine and mill.

The medical clinic will be equipped to treat general injuries and sicknesses, stabilize serious cases for medivac to off-site health facilities, and dispense drugs and medications. The clinic will have 12 beds, 6 of which will be in single rooms with adjoining washrooms. The facilities will include an emergency operating room, recovery room, x-ray room, two consulting/examination rooms, pharmacist’s office, doctor’s office, nurse’s office, nursing station, waiting room, and related support services, including a mortuary-type cold storage drawer. The clinic will initially be installed near the construction camp facility and relocated closer to the operations camp at the end of the construction phase of the project.

23.4.3 Port Site Infrastructure

The port site is located at Punta Rincón on the Caribbean Sea. Once construction is complete, it will become the main point of entry of supplies and equipment for the entire project site, including coal for the power plant, and the point of export for the copper concentrate.




On-Shore Facilities

These will include:

concentrate dewatering, storage, and handling facilities the power plant (to be constructed by an independent power producer) administration, warehouse, and shops complex container storage area and laydown fuel storage and distribution facilities 900-person construction camp (not including IPP camp) medical facilities helipad for emergency medical evacuations.

The administration, warehouse, and shops complex will house all the services, offices, and covered warehouse space required at the port site. The building will include a Panamanian customs office, medical facilities, and a garage for emergency and vehicles.

Much of the area within the port areas boundaries will be devoted to the storage of full containers unloaded from ships and empty containers returned from the plant site or other project locations and awaiting transfer to a ship.

Marine Facilities

These will include:

causeway access trestle shiploader service platform berthing dolphin and mooring arrangement fuel receiving system coal receiving system barge berth and ramp coastal freighter berth temporary construction quay.

The main berth is designed to handle ships in the design range from 30,000 dwt to 65,000 dwt for both copper concentrate export and coal import. The copper concentrate will be loaded into the ships by a radial shiploader with the capacity to cover up to three hatches without warping.

The barge and coastal vessel berth is designed to accommodate barges of up to 7,500 dwt and vessels of up to 10,000 dwt. The berth will allow receipt of all supplies for the mine, concentrator, and power plant. Supplies will generally be containerized in 20 ft containers and off-loaded by crane. The berth will be designed to accommodate roll-on/roll-off vessels.




A separate tug pen with sheltered mooring and refuelling facilities will be provided for tugs and line boats stationed at the terminal and Panamanian coast guard vessels.

23.4.4 Security Buildings

A security office/guard house and a station for the Panamanian police will be constructed at both the mine and port sites. The guard houses, manned by project security staff, will control the road access gates on the properties and be the centres of all site security functions for the facilities.

At the port site, an additional controlled security gate will be installed across the access road to the deep sea berth to provide security and controlled access to the offshore and marine facilities. In addition, the entire port site will be enclosed by a 2.1 m high fence.

The police stations will be used for all police force requirements for the general area and will be located outside the security perimeters of the site boundaries so that the police can provide local law enforcement functions and security support without violating any international, Homeland, or security/access requirements. One officer and eight personnel of other ranks will be assigned to each police station.

23.4.5 On-Site Roads

A network of general vehicle access roads around facilities and service roads to remote structures will be required at both the mine/plant and port sites; these will often be an extension of yard areas and will not always be delineated separately. The general access roads will be two-way and the service roads one-way with pullouts.

23.4.6 Access Roads to Site

Three access roads will serve the project development sites:

existing road from Penonomé via Coclecito and Molejón new Eastern Access Road new Coast Road from the port site to the mine site.

The existing road access is from the south, at a turnoff from the Pan-American Highway at Penonomé. This road runs northerly in the direction of the mine/plant site, bypassing the small community of Coclecito and continuing through the Molejón Sub-Concession.

MPSA will realign the road through the Molejón property around the mine workings during 2010 and early 2011 for use during the initial construction phase until the permanent Eastern Access Road is ready for use.

The Eastern Access Road will be roughly 12.6 km long, starting approximately 5.2 km west of the town of Coclecito. A bridge will be installed across Río Botija to allow the road to continue north to the southeast corner of the TMF, where it joins the Coast Road.




The Coast Road, including the reclaim pipeline corridor, will be approximately 30 km long and will be a private road to connect the project port site with the mine/plant site. The road will be suitable for the transport of regular and oversized freight and equipment needed for construction and operation of the mine/plant facilities. The three pipelines (concentrate, diesel fuel, filtrate return) will be buried in the shoulder of the road. Bridges will be required for crossings over Río Uvero and Río del Medio.

Measures to assist wildlife in crossing the roads will include the installation of netting across some sections for species living in the jungle canopy, local crossings over roadside ditches for ground-based species, and specially designed culverts for smaller wildlife.

23.5 Markets

23.5.1 Scope

The marketing plan for the Mina de Cobre Panamá project is based on updated metallurgical data, ongoing discussions and negotiations with potential customers, and analysis by independent consultants and Inmet Mining’s marketing group. All cost assumptions are based on consistent application of the principle that fuel and treatment charges are linked to the copper price used over the long term.

Inmet has estimated that the Mine de Cobre Panamá concentrate will be sold to smelters in Europe (35%) and Asia (65%). The forecast for average shipping rate for this study is $37/tonne.

23.5.2 Concentrate Production

The metallurgical results reported in the FEED Study indicate that the concentrates will be of good quality with no significant levels of deleterious constituents. A total of 27.3 million dry tonnes of copper concentrate at a grade of 28% Cu, containing 7.64 million tonnes of copper, is scheduled to be produced over the 30-year life of the mine. Production of molybdenum concentrate is expected to total 186,000 dry tonnes at a grade of 52%, containing 96,500 tonnes of molybdenum. The molybdenum concentrate will also contain 520 g/t of rhenium, which is potentially an added-value component.

The study is based on the production parameters for Years 1 to 30 as shown in Table 23-4.




Table 23-4: Mining Production Parameters

Copper Concentrate Molybdenum Concentrate

Year Ore Milled

(kt) Dry Weight

(tonnes) Grade (% Cu)

Copper Contained (tonnes)

Gold Contained

(oz)

Silver Contained

(oz) Dry Weight

(tonnes) Grade (% Mo)

Mo Contained (tonnes)

1 42,445 670,610 28 187,771 88,632 894,105 3,537 52 1,839

2 58,187 1,030,589 28 288,565 128,433 1,255,712 7,033 52 3,657

3 57,221 1,028,302 28 287,925 120,040 1,223,677 7,835 52 4,074

4 57,106 1,030,983 28 288,675 126,717 1,268,847 6,708 52 3,488

5 55,906 1,071,554 28 300,035 121,816 1,370,736 7,735 52 4,022

6 57,104 1,091,801 28 305,704 125,271 1,465,220 7,451 52 3,875

7 56,811 1,014,942 28 284,184 157,419 1,504,227 5,224 52 2,716

8 56,676 955,758 28 267,612 101,211 1,487,033 6,487 52 3,373

9 54,800 900,678 28 252,190 88,601 1,392,120 6,502 52 3,381

10 80,187 1,087,959 28 304,629 113,780 1,807,000 7,325 52 3,809

11 82,601 1,046,952 28 293,147 85,622 1,723,156 6,796 52 3,534

12 82,601 1,046,952 28 293,147 85,622 1,723,156 6,796 52 3,534

13 82,601 1,046,952 28 293,147 85,622 1,723,156 6,796 52 3,534

14 82,600 1,046,939 28 293,143 85,621 1,723,135 6,796 52 3,534

15 82,600 1,046,939 28 293,143 85,621 1,723,135 6,796 52 3,534

16 84,078 1,038,537 28 290,790 113,476 1,776,713 7,684 52 3,996

17 84,078 1,038,537 28 290,790 113,476 1,776,713 7,684 52 3,996

18 84,078 1,038,537 28 290,790 113,476 1,776,713 7,684 52 3,996

19 84,079 1,038,549 28 290,794 113,477 1,776,734 7,684 52 3,996

20 84,079 1,038,549 28 290,794 113,477 1,776,734 7,684 52 3,996

21 73,879 784,251 28 219,590 63,274 1,682,448 5,164 52 2,685

22 73,879 784,251 28 219,590 63,274 1,682,448 5,164 52 2,685

23 73,879 784,251 28 219,590 63,274 1,682,448 5,164 52 2,685

24 73,879 784,251 28 219,590 63,274 1,682,448 5,164 52 2,685

25 73,880 784,262 28 219,593 63,275 1,682,471 5,164 52 2,685

26 81,120 682,335 28 191,054 46,082 1,260,729 4,819 52 2,506

27 81,120 682,335 28 191,054 46,082 1,260,729 4,819 52 2,506

28 81,120 682,335 28 191,054 46,082 1,260,729 4,819 52 2,506

29 81,120 682,335 28 191,054 46,082 1,260,729 4,819 52 2,506

30 38,938 327,524 28 91,707 22,120 605,156 2,313 52 1,203

Life of Mine 2,142,652 27,288,749 28 7,640,850 2,690,230 45,228,358 185,648 52 96,537

2-20 1,367,393 19,640,010 28 5,499,203 2,078,779 30,273,918 134,701 52 70,045

21-30 732,814 6,978,129 28 1,953,876 522,819 14,060,335 47,410 52 24,653

23.5.3 Quality

The anticipated quality of the copper and molybdenum concentrates to be produced is based on extensive metallurgical testing carried out by Lakefield Research for the 1998 feasibility study and by G&T Metallurgical Services for the FEED study. The relevant parameters are presented in Table 23-5 and Table 23-6. Overall the concentrates are considered to be of good quality with no significant deleterious constituents.




Table 23-5: Analysis of Copper (Cu) Concentrate

Element Symbol Unit Typical Range

Copper Cu % 28.0 26-32

Silver Ag g/t 59 58-66

Gold Au g/t 3.3 1.9-6.3

Aluminum Al % 1.00 0.6-1.2

Antimony Sb g/t 31 10-80

Arsenic As g/t 20 10-25

Barium Ba % 0.012 0.001-0.02

Beryllium Be g/t <10 -

Bismuth Bi g/t 23 10-75

Cadmium Cd g/t 25 10-30

Calcium Ca % 0.62 0.2-0.9

Carbon C % 0.06 0.04-0.2

Chlorine Cl % <0.01 <0.01-0.02

Chromium Cr % 0.003 0.001-0.005

Cobalt Co g/t 62 44-84

Fluorine F g/t 150 70-320

Germanium Ge g/t <10 -

Indium In g/t <10 -

Iron Fe % 26.0 22-28.5

Lead Pb % 0.10 0.07-0.20

Magnesium Mg % 0.23 0.1-0.4

Manganese Mn % 0.02 0.01-0.02

Mercury Hg g/t <1 -

Molybdenum Mo % 0.06 0.01-0.3

Nickel Ni g/t 45 10-80

Phosphorus P g/t 200 50-200

Potassium K % 0.400 0.2-0.5

Selenium Se g/t 110 75-150

Silica SiO2 % 3.3 1-4

Sodium Na % 0.15 0.1-0.16

Sulphur S % 30.0 29-32

Tellerium Te g/t <10 -

Thallium Tl g/t <10 -

Tin Sn % 0.01 -

Titanium Ti % 0.03 0.03-0.06

Uranium U % <0.01 -

Vanadium V g/t 8 10-36

Zinc Zn % 0.30 0.2-0.6




Table 23-6: Analysis of Molybdenum (Mo) Concentrate

Element Symbol Unit Assay

Arsenic As g/t <10

Bismuth Bi g/t 136

Phosphorus P g/t <10

Calcium Ca % 0.75

Magnesium Mg % 0.094

Copper Cu % 1.81

Lead Pb % 0.042

Molybdenum Mo % 52.2

Sodium Na % 0.119

Rhenium Re g/t 520

23.6 Contracts and Agreements

No mining, smelting, refining, transportation, sales or hedging contracts have been finalized at the time of this writing. Any agreements will conform to industry standard practices.

23.7 Environmental Considerations

The Project will have an overall footprint of approximately 5,900 ha for the mine/plant site, the port site, and the connecting linear infrastructure. Of this area, approximately 2,500 ha of forest will be cleared in time for the start of operations and a total of 5,250 ha by the end of Mine life (30 years). The net disturbed area after reclamation during the operation and at closure will be 2,450 ha.

MPSA will create a mining operation of the highest standards that generates added value in a sustainable manner for the benefit of its shareholders, workers, neighbouring communities, and the country, and that protects the safety and health of its employees, the environment, and the surrounding communities. MPSA also intends to help establish an enduring economic base that will be viable beyond the life of the mine. To achieve this goal, MPSA will work with the communities in the region to develop sustainable infrastructure and economic activity through the mechanism of a Foundation that will bring together the communities, the regional government, and other representative bodies as partners for development, funded by a structured revenue-sharing arrangement based on MPSA’s mining activities. At the same time, MPSA recognizes that the project is located in a highly diverse ecosystem – part of the Mesoamerican Biological Corridor (MBC) – that is both sensitive and under threat from anthropogenic activity. MPSA will work with authorities to implement protected areas for the movement of fauna and to maintain habitat for threatened species, and will participate in education programs to help local residents understand the sensitivity of the ecosystem.

MPSA will also work with the communities and agencies to control to the extent possible project-induced in-migration and to create opportunities to alleviate pressures on natural ecosystems. The overall objective is to leave behind, at the end of the mine life, a sustainable




system that both supports local inhabitants in a reasonable manner and that maintains – and where possible enhances – the rich biodiversity of the region. MPSA has developed policies to demonstrate and communicate the company’s commitment to sustainability to all stakeholders.

MPSA is preparing an ESIA to international standards that identifies and addresses the potential sociological, socioeconomic, and environmental impacts of the Mina de Cobre Panamá project, both positive and negative. The ESIA conforms with the requirements of ANAM (Panamanian) and the International Finance Corporation (IFC) Performance Standards (PS) on Social and Environmental Sustainability for the assessment of a mining project of this magnitude. The assessment process included extensive consultations with stakeholders about the project, baseline studies of the local socioeconomic and natural environments, identification of concerns and opportunities arising from the field studies and consultations, and consideration of alternatives to and within the project to limit or mitigate potential negative impacts, and plans to address the remaining, residual project impacts.

MPSA’s community relations program started in 2007 and focused on the local communities to provide information on project plans and the alternatives being considered, and to listen to and understand community issues and concerns about the project. MPSA also engaged with external stakeholders, civil society, government agencies, and the media. Public ESIA-related consultations began in early 2008 and included two rounds of discussions with the same groups. In the second round, completed in August 2009, MPSA specifically discussed the expected impacts of the project and the proposed mitigation measures. Stakeholders again had an opportunity to question MPSA representatives and express their concerns.

Field and literature studies of the physical, biological, and socioeconomic environments were carried out to determine the existing state of these environments in the immediate area of the project (project development area, or PDA) and in the region surrounding the project (ecosystem region).

The project is located within one of the few remaining intact tracts of primary rainforest within the MBC. Both the Project development area (PDA) and the ecosystem region are relatively pristine. Habitat is intact in 89 percent of the PDA and in 92 percent of the ecosystem region.

The ESIA studies identified more than 850 species of flora, including 65 species “new to science” and 164 potential species of concern. More than 78 amphibians and reptile species, 262 bird species, 134 butterfly species, and 26 mammal species were found in the study area. In addition, four undescribed robber frog species considered “new to science” and endemic to the footprint were found.

Degraded areas include those of high human use for agriculture, placer mining, village residences along streams, and along exploration roads. Population in-migration and deforestation both appear to be increasing separately from the Project, reducing wildlife habitat and adversely affecting biodiversity.




Approximately 6,000 people reside within a 20 km radius of the proposed mine, mainly in small communities varying in size from 100 to 600 people. Latino and indigenous communities are both present. Most of these communities have no infrastructure, only rudimentary health facilities and basic schooling. Average income ranges from US$300 to US$400 per month per family, with many families dependent on subsistence agriculture or activities such as artisanal mining. Water resources for human use are abundant, although drinking water quality is poor, with high levels of fecal coliform in many potable water sources.

Social conditions throughout the area are changing in response to perceived opportunities associated with other mining projects and artisanal gold mining activities, and as a result of in-migration of indigenous people to the forest. Increasing deforestation resulting from agricultural, mining, hunting, and fishing pressures is leading to local scarcities of some plant and animal species.

A human health and ecological risk assessment (HHERA) found that most combinations of existing receptors and contaminants of potential concern found negligible or low risks to human health and the ecology.

A valued ecosystem components (VEC) approach was used to assess the potential impacts of the project. The selected VECs incorporate both the natural and human environments and in large part reflect the issues and concerns that emerged during the stakeholder consultations.

Direct loss of habitat will primarily result from clearing associated with the project; indirect loss of habitat will primarily result from the in-migration of people to the area, some of which is occurring as a baseline condition and some of which will be induced by project development. This habitat loss will negatively affect the biodiversity of the regional ecosystem as well as the movement of fauna through established forest corridors. Habitat loss and its direct effects on flora and fauna, and particularly on identified species of concern, is the largest biophysical impact of the project on its surroundings.

Removal of the forest cover and reduced access to the natural resources of the forest will have negative effects on the populations that depend on these resources for their livelihood. Examples of such resources are construction materials (wood), medicinal plants, fauna, fish, and, in the port area, marine resources. Further deterioration of natural resources could result from long-term project-induced in-migration of people into the area.

Some individuals and families will be forced to relocate because they are located within the proposed project footprint; these people will be assisted through MPSA’s resettlement plan which will comply with IFC PS 5 and 7.

The project will, however, result in the generation of employment for local communities and of higher income and revenue through new job opportunities, procurement of goods and services, and payment of taxes and royalties – a positive impact.




Measures to limit and/or mitigate impacts on the natural and socioeconomic environments have been incorporated into project planning and design. The residual impacts of the project are those negative or positive effects that remain after all mitigations have been implemented. The principal residual biological impact will be the reduction of biodiversity in the region resulting from development within the project footprint. Further mitigation will depend on the success of the project reclamation program designed to restore secondary growth (initially) to the PDA.

The other principal residual impacts will be on the social and economic structures of the region. Whether these are positive or negative will depend on the success of the efforts of the Foundation and other MPSA initiatives to offset social issues and establish a sustainable economic system that extends beyond a dependence on mining.

Further residual impacts will relate to the creation of four large bodies of standing water (the three mine pits and the tailings pond), possible long-term effects on water quality that may persist despite of water treatment, and the long-term stability of the tailings management facility embankments. In all cases, MPSA is committed to ensuring that the mitigation actions remain as effective as possible through the mine life and after closure.

MPSA has developed mitigation plans and actions based on evolving international best practice, particularly in the areas of biodiversity offsets and compensation and of community development. MPSA has developed an environmental and social management system to limit and/or mitigate negative effects and provide positive benefits to the local and national economy.

At the time of this writing, no reclamation/remediation bonds have been posted for the proposed project development.

23.8 Taxes and Royalties

Law No. 9, 1997, which governs the Project Concession, provides for payment of a royalty of 2% of “Negotiable Gross Production” (defined as “the gross amount received from the buyer due to the sale (of concentrates) after deducting all smelting costs, penalties and other deductions, and after deducting all transportation costs and insurance…incurred in their transfer from the mine to the smelter”) as well as payments of 3 Panamanian Balboas (US$3.00) per hectare per year of Concession area.

In addition, Law 9 provides certain fiscal incentives including:

1. “Exoneration for the Company, its Affiliates, contractors and subcontractors of any import tax or duty, contribution, charge, consular fee, lien, duty or another tax or contribution, or of any name or class that fall [are levied] on the introduction and import of equipment, machinery, materials, parts, diesel and Bunker C and other petroleum derivatives”

2. “Income tax exoneration applicable to remittances or transfers abroad, made to pay commissions, loans, royalties, returns, charges for professional advice or administration incurred outside the national territory”




3. “Excepting only the respective mining royalties and royalties, as long as the Company has not finished repaying the debt which the Company or its Affiliates acquire for construction and development of [the Project], the Company and its Affiliates shall be totally exempt from payment of any type of tax, fee, duty, charge, lien, contribution or tribute that may be levied due to any reason in relation to the development of THE PROJECT, except municipal taxes.”

For the purposes for this study it is assumed the term of the debt repayment schedule will be 11 years. These benefits and charges are incorporated into the financial model.

All quotations in this section are from an English translation of Law 9.

23.9 Capital Costs

The total estimated cost to design and build the Mina de Cobre Panamá 150,000 t/d project described in this report is US$4,320 million, including an Owner-provided mining fleet and self-performed preproduction development. This amount covers the direct field costs of executing the project, plus the Owner’s indirect costs associated with design, construction, and commissioning. The estimate is summarized in Table 23-7.

The capital cost estimate is based on an EPCM approach. Mid-way into the study, MPSA asked that AMEC change to an EPC-type approach, and allowances were made to capture costs associated with an EPC contract, as follows:

EPC contractor procured material – overheads and fees Construction equipment rental – overheads and fees Subcontractor management – overheads and fees.

AMEC was responsible for developing the costs for its scope of work and for assimilating costs provided by other project participants into an overall project capital cost estimate, as follows:

Sandwell Engineering Inc ................... marine facilities design, quantities, and estimates Pipeline Systems International (PSI) ...... concentrate, water return, and diesel pipelines Estudios Electricos ............................. 230 kV power line and Llano Sánchez substation MPSA ............................................. mining preproduction development and mining fleet MPSA ........................................................................................................ Owner’s costs

With the exception of initial earthworks, the estimate does not include capital costs for the power generating plant and associated facilities being constructed by others.

The scope of the project evolved over the course of the FEED Study. Changes made after December 2009 included relocating the construction and permanent camps at the plant site, adding the Eastern Access Road, and modifying the plant site layout to allow for the future third processing train. Where engineering assessment was performed for the changes, the cost impact has been included in the capital cost build-up. Where engineering assessment was not completed to a full FEED level, allowances were made and added to the appropriate cost area.




All costs are expressed in second quarter 2009 (Q2) US dollars. No allowance has been included for escalation, interest, financing fees, taxes, duties, or working capital during construction. The level of accuracy for the estimate is ±15% of estimated final costs, as per AACE Class 3 definition.




Table 23-7: Summary of Capital Costs

WBS Area (US$000)

1000 Mining 1110 Mine Preproduction Development 191,673 1130 Mine Dewatering and Drainage 25,849 1140 Mine Mobile Equipment 287,510 1150 Blasting Agent Storage 760 1160 Emulsion Plant 4,111 Total Mining 509,903

2000 Process Plant 2100 Primary Crushing and Conveying 120,380 2200 Concentrator 471,839 2300 Tailings System 207,210 Total Process Plant 799,429

3000 Site and Services 3100 Site Preparation and Civil Works 293,325 3200 Plant Site Facilities 75,146 3300 Electrical and Heating Services 94,220 3400 Water Services 22,095 3500 Fuel 4,306 3600 Waste Handling 9,755 3700 Mobile Equipment 43,841 3800 Plant Site Construction Equipment 54,558 Total Site and Services 597,247

6000 Port Site Facilities 6100 Port Concentrate Handling 80,485 6200 Port Site Development 95,727 6300 Port Site Facilities and Services 107,668 6395 Port Site Substation and Power Distribution 10,910 6410 Coal Receiving Hopper 17,401 6500 Port Site – Construction Equipment 7,760 Total Port Site Facilities 319,951

Total Directs 2,226,530

8000 Construction Indirects 8100 Construction Services 592,385 8200 Mobilization and Shipping 139,326

9000 Other Indirects 9100 Start-up and Commissioning 68,069 9200 Owner’s Costs 364,381 9300 EPCM + EPC Allowance 472,002 9345 Transmission Line – EPC 4,786

Total Indirects 1,640,950

9500 Contingency 452,896 9530 Escalation (excluded) - Working Capital (excluded) -

Total Capital Cost 4,320,375




Sustaining capital, not covered in Table 23-7, includes costs for the installation of the Colina crushing station and associated materials handling infrastructure, replacement of the mine and plant mobile equipment, and the installation of the third grinding line to increase production in Year 10. The cost estimate for these major expansions and equipment is $1,116 million.

Sustaining costs also include typical items such as ongoing tailings dam construction, water management and treatment facilities, electrical equipment, and community support. Total life-of-mine sustaining capital for these items, plus the mobile equipment and plant expansions mentioned above, is estimated at $1,723 million.

23.10 Operating Costs

The operating cost estimate (opex) has been prepared as an annual cost for the project from plant start-up to mine closure. The life of the project is 30 years at a nominal processing plant throughput rate of 150,000 t/d (54.8 Mt/a) for the first nine years, followed by an increase to a nominal throughput of 225,000 t/d (82.1 Mt/a) from Year 10 onward.

The operating cost estimate is expressed in constant second quarter 2009 (Q2 2009) U.S. dollars with no allowances for escalation or fluctuation in exchange rates. Costs incurred before plant start-up on Q4 April 2015 are treated as capital expenditures (capex).

The operating costs are grouped into four cost centres:

Open pit mining Processing Site services General and administration (G&A).

The average costs for the project over the mine life are shown in Table 23-8.

Table 23-8: Summary of Operating Cost Estimate ($/t milled)

Cost Centre Labour Material Power Other Total

Open Pit Mining 0.16 1.69 0.06 0.22 2.14

Processing 0.11 1.44 2.12 0.05 3.72

Site Services 0.12 0.18 0.06 0.38 0.73

G&A 0.09 0.00 0.02 0.52 0.64

Total 0.48 3.31 2.26 1.18 7.23

The total operating cost is $15,469 million over the LOM for a milled feed of 2,143 Mt. The overall unit operating cost is $7.23/t of milled ore.




23.11 Economic Analysis

Summary

Engineering studies have demonstrated the technical feasibility of producing significant quantities of copper, molybdenum, silver, and gold from the Mina de Cobre Panamá project. The economic viability of the project has been evaluated by using a combination of pre-tax and after-tax cash flow analyses, based on the engineering studies and cost estimates discussed herein. Under the metal price assumptions shown in Table 23-9, and using a discount rate of 8%, the pre-tax project net present value (NPV) for the base case is $1,661 million, and the internal rate of return (IRR) is 12.6%. The after-tax NPV is $1,536 million with an IRR of 12.4%. The cumulative pre-tax undiscounted cash flow value for the project is $12,873 million and the payback period is 5.9 years.

Another case (levered) assuming $2.16 billion of debt financing, representing 50% of the preproduction capital, was evaluated. This resulted in a project after-tax IRR of 15.1% – a net improvement of 2.7% over the un-levered base case.

Table 23-9: Long-Term Metal Price Assumptions

Metal Unit Price

Cu US$/lb 2.10

Au US$/oz 885.00

Mo US$/lb 13.00

Ag US$/oz 13.50

All monetary amounts are presented in United States dollars (US$).

For the sake of discounting, cash flows are assumed to occur at the end of each period. All cash flows are discounted to the beginning of Q4 2010.

Base Case

Table 23-10 summarizes the key project financials for the un-levered base case.

Sensitivity analyses for the project NPV @ 8% and the IRR were performed on a range of metal prices (Cu, Au, Ag, Mo), from -30% to +30%, as well as changes to capital costs, operating costs, steel prices (grinding media), process reagent prices, and diesel fuel prices. The results are shown in Figure 23-6 and Figure 23-7.

The project is most sensitive to changes in metal prices (as well as recovery and head grade), less so to capital and operating cost changes, and least sensitive to variations in individual commodity pricing for steel, reagents, and diesel. The dramatic effect of price change can be observed in Table 23-11 .




Table 23-10: Summary of Key Financials (base case)

Valuation indicator Unit LOM

Pre Tax

Cumulative net cash flow US$M 12,873

NPV (8%) US$M 1,661

Payback period Years 5.9

After Tax

Cumulative net cash flow US$M 12,004

NPV (8%) US$M 1,536

Payback period Years 5.9

IRR and Tax Rate

Effective tax rate % 4.57

IRR before tax % 12.6

IRR after tax % 12.4

Figure 23-6: Sensitivity of After-Tax NPV @ 8%

Sensitivity of After Tax NPV @ 8%

(2,000)

(1,000)

0

1,000

2,000

3,000

4,000

5,000

-40% -30% -20% -10% 0% 10% 20% 30% 40%

Change in Factor

NP

V @

8%

($U

S m

illio

n)

Capital expenditure Operating expenditure M etal price Steel price Reagents Copper Diesel Fuel




Figure 23-7: Sensitivity Spider Graph for After-Tax IRR

Sensitivity of After Tax IRR

0%

2%

4%

6%

8%

10%

12%

14%

16%

18%

20%

-40% -30% -20% -10% 0% 10% 20% 30% 40%

Change in Factor

IRR

(%

)

Capital expenditure Operating expenditure M etal price Steel price Reagents Copper Diesel Fuel

Table 23-11: Impact of Metal Price Change

Item Unit Base + 5% + 10% + 14% + 19% + 24% + 29% + 33%

Copper US$/lb 2.10 2.20 2.30 2.40 2.50 2.60 2.70 2.80

Gold US$/oz 885 927 969 1,011 1,054 1,096 1,138 1,180

Molybdenum US$/lb 13.00 13.62 14.24 14.86 15.48 16.10 16.71 17.33

Silver US$/oz 13.50 14.14 14.79 15.43 16.07 16.71 17.36 18.00

Un-levered Base Case After-Tax Financial Results

Cumulative Net C.F. (US$M) 12,004 13,342 14,682 16,023 17,361 18,696 20,032 21,368

Net Present Value 8% (US$M) 1,536 1,914 2,292 2,670 3,048 3,426 3,803 4,181

Internal Rate of Return (%) 12.4 13.4 14.3 15.2 16.0 16.9 17.7 18.5

Levered Case

The levered case assumed the following:

debt financing interest rate of 7%

$2.16 billion of debt representing 50% of preproduction capital

debt repayment term of 10 years with a one-year repayment holiday.

The financial results of the levered case at various pricing scenarios are shown in Table 23-12.




Table 23-12: Levered Case Financial Results

Item Unit Base + 5% + 10% + 14% + 19% + 24% + 29% + 33%

Copper US$/lb 2.10 2.20 2.30 2.40 2.50 2.60 2.70 2.80

Gold US$/oz 885 927 969 1,011 1,054 1,096 1,138 1,180

Molybdenum US$/lb 13.00 13.62 14.24 14.86 15.48 16.10 16.71 17.33

Silver US$/oz 13.50 14.14 14.79 15.43 16.07 16.71 17.36 18.00

Levered After-Tax Financial Results

Cumulative Net C.F. (US$M) 10,949 12,287 13,627 14,968 16,306 17,641 18,977 20,313

Net Present Value 8% (US$M) 1,697 2,075 2,453 2,831 3,209 3,586 3,964 4,342

Internal Rate of Return (%) 15.1 16.5 17.8 19.1 20.4 21.6 22.7 23.9

Performance Statistics

Inmet cost metrics are summarized in

Table 23-13 for the base case and in Table 23-14 for the levered case. The metric items are defined as follows:

Cash costs are the sum of on-site costs, off-site costs, NSR royalty, non-income taxes, and by-product credits divided by recovered copper.

Break-even cash costs are cash costs plus sustaining capital divided by recovered copper.

Financed break-even cash costs are break-even cash costs plus interest expense divided by recovered copper.

Total costs are cash costs plus interest expense divided by recovered copper plus life-of-mine capital divided by life-of-mine recovered copper.

Table 23-13: Summary of Base Case Performance Statistics

Cash Costs Units Year 1 Year 2 Year 3 Avg. Y2-16 LOM

Inmet-Defined Cost Metrics

Cash costs US$/lb 1.01 0.67 0.66 0.78 0.90

Break-even cash costs US$/lb 1.01 0.78 0.86 0.92 1.00

Financed break-even cash costs US$/lb 1.01 0.78 0.86 0.92 1.00

Total costs US$/lb 1.37 1.03 1.01 1.14 1.26

Table 23-14: Summary of Levered Case Performance Statistics

Cash Costs Units Year 1 Year 2 Year 3 Avg. Y2-16 LOM

Inmet-Defined Cost Metrics

Cash costs US$/lbs 1.01 0.67 0.66 0.78 0.90

Break-even cash costs US$/lbs 1.01 0.78 0.86 0.92 1.00

Financed break-even cash costs US$/lbs 1.40 1.02 1.07 1.00 1.06

Total costs US$/lbs 1.75 1.27 1.23 1.23 1.31




23.12 Payback

Payback for the project is 5.9 years.

23.13 Mine Life

The mine life of the Mina de Cobre Panamá project is 29.5 years after initial mill startup.

mina de cobre panamÁ project panamÁ ni 43-101 …mina de cobre panamÁ project, panamÁ ni 43-101...

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