report on feasibility study update ruby … study update ruby creek project, ... table 11.2 drill...

REPORT ON

Submitted to: Adanac Molybdenum Corporation

#200-2055-152nd Street Surrey, BC V4A 4N7

Prepared by: Rick Alexander, P.Eng

December, 2007

FEASIBILITY STUDY UPDATE RUBY CREEK PROJECT,

NORTHERN BRITISH COLUMBIA, CANADA

Ruby Creek Feasibility Study Update - i - December, 2007

TABLE OF CONTENTS SECTION PAGE

TABLE OF CONTENTS.......................................................................................... I 1.0 SUMMARY.................................................................................................. 1 2.0 INTRODUCTION....................................................................................... 11 3.0 RELIANCE ON OTHER EXPERTS.......................................................... 13 4.0 PROPERTY DESCRIPTION AND LOCATION ........................................ 14 5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE

AND PHYSIOGRAPHY............................................................................. 18 6.0 HISTORY .................................................................................................. 20 7.0 GEOLOGICAL SETTING.......................................................................... 24

7.1 Regional Scale ......................................................................................24 7.2 Local and Property Scale ......................................................................26

8.0 DEPOSIT TYPES...................................................................................... 29 9.0 MINERALIZATION.................................................................................... 30 10.0 EXPLORATION......................................................................................... 32

10.1 ADANAC 2004.......................................................................................32 10.2 ADANAC 2005.......................................................................................32 10.3 ADANAC 2006.......................................................................................33 10.4 2007 Drilling Program............................................................................35

11.0 DRILLING.................................................................................................. 36 12.0 SAMPLE METHOD AND APPROACH..................................................... 40 13.0 SAMPLE PREPARATION, ANALYSES AND SECURITY ....................... 42

13.1 Field Sample Preparation Procedures...................................................42 13.1.1 Adanac 2004, 2005 and 2006....................................................42 13.1.2 Prior to Adanac ..........................................................................43

13.2 LABORATORY SAMPLE PREPARATION PROCEDURES .................44 13.2.1 2004, 2005 and 2006 ADANAC.................................................44 13.2.2 1979 — 1980 Placer ..................................................................44 13.2.3 Pre-1979 Kerr Addison ..............................................................44

13.3 Analytical Procedures............................................................................45 13.3.1 Adanac 2004, 2005 and 2006....................................................45 13.3.2 1979 — 1980 Placer ..................................................................46 13.3.3 Pre-1979 Kerr Addison ..............................................................46

13.4 Quality Assurance and Quality Control..................................................46 13.4.1 Adanac Procedures ...................................................................46 13.4.2 QA/QC Results –2006 Samples ................................................47

14.0 DATA VERIFICATION .............................................................................. 49 15.0 ADJACENT PROPERTIES....................................................................... 50

Ruby Creek Feasibility Study Update - ii - December, 2007

16.0 MINERAL PROCESSING AND METALLURGICAL TESTING................ 51 16.1 Mineral Processing ................................................................................51 16.2 Metallurgical Testing..............................................................................52

16.2.1 Kerr Addison’s Pilot Plant (Britton 1969 – 1971) .......................53 16.2.2 SGS-MinnovEX Test Work ........................................................53 16.2.3 B.C. Mining Research Limited Test Work..................................54 16.2.4 G&T Test Work ..........................................................................54

17.0 MINERAL RESOURCE AND MINERAL RESERVE ESTIMATE............. 56 17.1 The Database ........................................................................................57 17.2 The Geological Model............................................................................58 17.3 Wireframe Validation .............................................................................61 17.4 Data Preparation and Compositing .......................................................62 17.5 Declustering...........................................................................................63 17.6 Spatial Trend Analysis...........................................................................63 17.7 High-Grade Treatment...........................................................................63 17.8 Variogram Analysis................................................................................64

17.8.1 Variography Objectives and Approach ......................................64 17.8.2 Summary of Variography Parameters .......................................65

17.9 Update Block Model Parameters...........................................................65 17.10 Grade Interpolation................................................................................67

17.10.1 Grade Interpolation Methods and Objectives.................67 17.10.2 Ordinary Kriging Plan.....................................................67

17.11 Density Assignment...............................................................................68 17.12 Mineral Resource Classification ............................................................69 17.13 Mineral Resource Summary ..................................................................70 17.14 Mineral Reserve Estimate and Mine Design .........................................72

18.0 OTHER RELEVANT DATA AND INFORMATION ................................... 82 18.1 Tailings Facilities, Waste Rock Dumps and Site Water Management...82

18.1.1 Tailings Characterization ...........................................................82 18.1.2 Tailings Facility ..........................................................................82 18.1.3 Waste Rock Dumps ...................................................................84 18.1.4 Site Water Management Plan....................................................85 18.1.5 Operating and Monitoring Controls............................................86

18.2 Mine Closure Plan .................................................................................87 18.2.1 Tailings Facility ..........................................................................87 18.2.2 Waste Dump/East Ruby ............................................................87 18.2.3 Open Pit.....................................................................................88 18.2.4 Temporary Mine Closure ...........................................................88

18.3 Environmental Considerations...............................................................89 18.4 Capital Costs .........................................................................................91

Ruby Creek Feasibility Study Update - iii - December, 2007

18.5 Operating Costs.....................................................................................94 18.6 Market....................................................................................................96

18.6.1 Supply Fundamentals ................................................................97 18.6.2 Supply Outlook ..........................................................................97 18.6.3 Demand Fundamentals .............................................................99 18.6.4 Molybdenum in Steel .................................................................99 18.6.5 Other Molybdenum Applications..............................................100 18.6.6 Substitutes ...............................................................................102 18.6.7 End-Use Industry Analysis.......................................................102 18.6.8 Price Outlook ...........................................................................103

18.7 Economic Model ..................................................................................105 18.7.1 Net Present Value and Internal Rate of Return Summary.......105 18.7.2 Sensitivity Analysis ..................................................................106 18.7.3 Summary of Results ................................................................107

19.0 INTERPRETATIONS AND CONCLUSIONS.......................................... 109 20.0 RECOMMENDATIONS........................................................................... 109 21.0 REFERENCES........................................................................................ 110 22.0 DATE AND SIGNATURE PAGE............................................................. 113 LIST OF TABLES Table 1.1 February 22, 2007 Mineral Resource Estimate Table 1.2 2007 Updated Mineral Reserves Table 1.3 Capital Cost Summary Table 1.4 Operating Cost Summary (First Five Years of Full Production) Table 1.5 Operating Cost Summary (After Five Years of Full Production) Table 1.6 Molybdenum Price Outlook Table 1.7 Production and Operating Costs (First Four Years of Production) Table 1.8 IRR and NPV Summary Results Table 4.1 List of Mineral Claims Table 6.1 February 8, 2006 Mineral Resource Estimate Table 11.1 Drilling History of the Ruby Creek Deposit (Dec 31, 2006) Table 11.2 Drill Holes in the 2006 Ruby Creek Datamine Database Table 17.1 Drill Hole Data in the July 23, 2007 Mineral Resource Update Table 17.2 2007 Ruby Creek Molybdenum Project 2D and 3D Geometries Table 17.3 Primary and Secondary Lithological Units Table 17.4 High Grade Thresholds for % Mo by Zone Table 17.5 Variography Parameters for Zones 6, 60 and 7 Table 17.6 Block Model Dimensions for Ruby Creek Resource Model Table 17.7 Kriging Plan Parameters

Ruby Creek Feasibility Study Update - iv - December, 2007

Table 17.8 Rock Type Bulk Density Assigned to the Block Model Table 17.9 February 22, 2007 Mineral Resource Estimate Table 17.10 November 22, 2007 Mineral Reserve Estimate Table 17.11 Production Schedule Table 17.12 Mine Department Personnel Table 18.1 Capital Cost Summary Table 18.2 Operating Cost Summary (First Five Years of Full Production) Table 18.3 Operating Cost Summary, (After Five Years of Full Production ) Table 18.4 Molybdenum Price Outlook Table 18.5 IRR and NPV Summary Results LIST OF FIGURES Figure 4-1 Location Map Figure 4-2 Mineral Claims Map1

Figure 7-1 Regional Geology Map of the Atlin Area Figure 10-1 Plan View of 2006 Drill Hole Collars Figure 11-1 Plan View of 2007 Drill Hole Collars Figure 13-1 Adanac Field Sample Preparation Procedures Figure 13-2 Kerr Addison’s Laboratory Sample Preparation Procedure Figure 16-1 Simplified Flowsheet Figure 17-1 Section View of 3D Mineralization Zones and Rock Types Figure 17-2 2006 Global Distribution of Raw Sample Lengths (Palmer, 2006) Figure 17-3 Plan View of Block Model Geometry Figure 17-4 Summary Schedule Figure 17-5 Ultimate Pit Design Figure 17-6 Phase Pushback Sequence in Plan Figure 17-7 Phase Pushback Sequence in Section Figure 18-1 Real Molybdenum Prices and World Supply and Demand Balance LIST OF APPENDICES Appendix A Economic Model Appendix B Certificate of Qualification

Ruby Creek Feasibility Study Update - 1 - December, 2007

1.0 SUMMARY

The Ruby Creek Molybdenum Project (“the Project”) is a large porphyry molybdenite deposit located approximately 24 km northeast of Atlin, British Columbia, which potentially would include an open pit mine and ore processing facility.

This report was prepared by Rick Alexander, P.Eng (Alexander) at the request of Adanac Molybdenum Corporation (“Adanac”), in order to prepare a feasibility study to incorporate all new information on the project including current resource and reserve estimates and to outline the standards of disclosure required for mineral projects under National Instrument (“NI”) 43-101. The work has included a resource estimate, mine design and the development of a metallurgical process. This NI 43-101 compliant Technical Report has been based on work by Adanac’s in house professional staff, G&T, SGS MinnovEX, Wardrop, Golder, Klohn Crippen, and CPM Group.

Discovered in 1905, it has been explored on several occasions (between 1968 to 1981) but failed to advance into production on account of molybdenum prices holding in the US$2.00 - US$4.00 per pound molybdenum range during that period. The present owners, Adanac, acquired a 100% interest in the Project, (no royalties) in 2002 through staking 189 units, which covers the upper southwest part of Ruby Creek Valley and much of the adjacent Boulder Creek Valley. Six mining claims were converted into a Mining Lease for development on March 27, 2007.

In April 2005, Adanac commissioned a team of engineering consultants to complete the component studies of a Preliminary Feasibility Study (as defined by NI 43-101) on the Project. Based on the results of that study, it was recommended that the project proceed to the Feasibility Study level. As such, in September 2005, the following engineering consulting companies were commissioned to complete the component studies for a Feasibility Study (as defined by NI 43-101):

• Wardrop Engineering Inc. (Wardrop):

- Feasibility Study Report, 2006 (“the 2006 Feasibility Study”)

• Golder Associates Ltd. (Golder):

- Ruby Creek Molybdenum Project - Mining Feasibility Study, 2006

- Technical Report - Mineral Resource Estimate Ruby Creek Molybdenum Project, February 8, 2006

- Pit Slope Stability Considerations for the Ruby Creek Project, 2006


• Wardrop Engineering Inc.:

- Feasibility, Process and Infrastructure Design and Cost Estimate, Ruby Creek, 2006

• Klohn Crippen Berger Consultants Ltd. (Klohn Crippen):

- Feasibility Design of Tailings Facility, Waste Dumps and Site Water Management, Ruby Creek Project, 2006

Alexander has revised the original 2006 Feasibility Study Report to incorporate new and updated information provided by the authors of the component studies, which include:

• Adanac:

- Internal reports prepared by professional engineers employed by Adanac and reviewed by independent qualified persons (QP) to confirm the work was performed in accordance with good and accepted engineering practice.

• Klohn Crippen:

- Site Water Management Design Report, February 2007;

- Tailing Facility Detail Design Report, April 2007

• G&T Metallurgical Services Ltd (G&T):

- An Assessment of Metallurgical Response, Ruby Creek Project, December 2006

• CPM Group (CPM Group):

- Sustainability of Recent Molybdenum Prices, a Molybdenum Industry Analysis, October 2007 (“the October 2007 CPM Marketing Report”)

• Golder:

- Mineral Resource Estimate Update, Ruby Creek Molybdenum Project, July 2007

This revision to the 2006 Feasibility Study is entitled “Feasibility Study Update, Ruby Creek Project, Northern British Columbia, Canada” (“the Feasibility Study Update”).

The Project is planned as an open pit mining operation with on-site ore beneficiation. The mineral resource is a porphyry molybdenum deposit in multiple phases of felsic intrusions. Based on the ore characteristics and reserve magnitude, the mill has been designed to operate at an average rate of approximately 23,000 metric tonnes per day for a life of 21 years. The site has no developed infrastructure with the exception of a single lane access road.


Mining and processing are scheduled on a 24 hour per day, 7 days per week schedule. The Project will be operated as a fly in/fly out camp accommodation program, with crews working on a two week turn-around basis. The operating management and labour positions are expected to peak at 250 employees. Whereas it is located at an elevation of some 1,400 metres above sea level, the Project is also easily accessible by road from Atlin.

Golder completed an NI 43-101 compliant Technical Report entitled “Technical Report - Mineral Resource Estimate, Ruby Creek Project, British Columbia” and dated February 8, 2006 (“the February 8, 2006 Golder Technical Report”), which describes the geology and this was the second time that Adanac reported resources to the public. Golder then completed an update to the February 8, 2006 Golder Technical Report estimate to incorporate the results of the 2006 drilling program. An updated Mineral Resource Estimate for the Ruby Creek Property was provided to Adanac on February 22, 2007 as a letter report (“the February 22 Mineral Resource Estimate”), which was reproduced in Golder’s report entitled “Mineral Resource Estimate Update, Ruby Creek Molybdenum Project” and dated July 23, 2007 (“the July 23, 2007 Mineral Resource Update”). The July 23, 2007 Mineral Resource Update does not include any drilling information collected during the 2007 drilling exploration program for the Ruby Creek Project. The July 23, 2007 Mineral Resource Update was used in the 2007 Feasibility Study Update .

The February 22, 2007 Mineral Resource Estimate, as reported in the July 23, 2007 Mineral Resource Update, is tabulated by cut-off grades from 0.02 to 0.1 % molybdenum (%Mo) for measured, indicated and inferred mineral resource categories and is summarized in Table 1.1.


Table 1.1 February 22, 2007 Mineral Resource Estimate

Resource Category Cut-off (%Mo) Tonnage %Mo Mo lb

0.020 55,831,000 0.068 83,698,000 0.030 54,300,000 0.069 82,600,000 0.040 49,106,000 0.073 79,029,000 0.050 41,389,000 0.078 71,172,000 0.060 30,151,000 0.086 57,165,000 0.070 21,909,000 0.094 45,403,000 0.080 14,556,000 0.104 33,374,000 0.090 10,411,000 0.112 25,706,000

Measured

0.100 6,612,500 0.122 17,785,000 0.020 387,278,000 0.042 358,593,000 0.030 238,954,000 0.052 273,935,000 0.040 163,801,000 0.060 216,669,000 0.050 109,444,000 0.067 161,658,000 0.060 61,471,000 0.077 104,350,000 0.070 37,664,000 0.084 69,749,000 0.080 18,813,000 0.094 38,987,000 0.090 9,848,100 0.102 22,145,000

Indicated

0.100 4,286,900 0.113 10,680,000 0.020 443,108,000 0.045 442,290,000 0.030 293,254,000 0.055 356,535,000 0.040 212,907,000 0.063 295,699,000 0.050 150,834,000 0.070 232,831,000 0.060 91,621,000 0.080 161,513,000 0.070 59,573,000 0.088 115,151,000 0.080 33,369,000 0.098 72,360,000 0.090 20,259,000 0.107 47,851,000

Measured +

Indicated

0.100 10,899,000 0.118 28,464,000 0.020 135,737,000 0.032 95,759,000 0.030 48,456,000 0.045 48,072,000 0.040 24,973,000 0.054 29,730,000 0.050 11,631,000 0.064 16,411,000 0.060 5,194,300 0.077 8,817,500 0.070 2,626,200 0.089 5,152,900 0.080 1,239,700 0.103 2,815,000 0.090 821,310 0.113 2,046,000

Inferred

0.100 493,790 0.125 1,360,800 Cut-off %Mo grades were classified as greater than or equal to and range from 0.01 to 0.10 in increments of 0.01. The mineral resource estimate has been completed in accordance with CIM standards of Estimation of Mineral Resources and Mineral Reserves.


The combined Measured and Indicated Mineral Resource estimate using a 0.04 %Mo cut-off is 212,907,000 metric tonnes (T) with a grade of 0.063 %Mo and 295,699,000 pounds of molybdenum. Comparing the February 22, 2007 Mineral Resource Estimate to the previous estimate in the February 8, 2006 Golder Technical Report shows an increase of some 6,532,000 tonnes of measured and indicated resources and an increase of 10,095,000 pounds of contained molybdenum, with no significant change in grade. The changes from the February 8, 2006 Mineral Resource Estimate is attributed to the 2006 inclined drilling program (with respect to grade and tonnage) and re-interpretation of the 2006 mineralized geometries (affecting tonnage only).

An updated Mineral Reserve Estimate was developed based on the February 22, 2007 Mineral Resource Estimate and on a new and updated mine design, using a 0.04 %Mo mining grade cut-off and a 0.03 %Mo milling grade cut-off. This is provided in Table 1.2.

Table 1.2 2007 Updated Mineral Reserve

Ore to Mill Stockpile Ore Proven Probable Proven Probable Total Tonnes %Mo Tonnes %Mo Tonnes %Mo Tonnes %Mo Tonnes %Mo

Phase 1 19,455,000 0.089 3,065,000 0.082 6,081,000 0.049 2,608,000 0.042 31,209,000 0.077 Phase 2 3,903,000 0.070 6,819,000 0.075 4,996,000 0.049 7,519,000 0.046 23,237,000 0.059 Phase 3 20,250,000 0.056 48,463,000 0.050 185,000 0.027 2,893,000 0.026 71,791,000 0.051 Phase 4 271,000 0.049 29,166,000 0.056 12,000 0.027 1,999,000 0.026 31,448,000 0.054 Total 43,879,000 0.072 87,513,000 0.055 11,274,000 0.049 15,019,000 0.039 157,685,000 0.058

Upon completion of open pit mining of Phases 1 to 4 an estimate of 157,564,000 tonnes with a grade of 0.058 %Mo is extracted based on the updated mine design.

In their 2006 work, Wardrop made use of detailed pilot plant information developed by two of the previous operators of the property (Kerr Addison Mines Limited and Placer Development Limited), and additional documentation provided by Adanac and its consultants. The information available from the earlier work (1969 to 1981) was updated and supported by supplementary data developed in years 2004 and 2005. SGS-MinnovEX Technologies Inc. (MinnovEX) was contracted by Adanac to develop comminution and flotation studies to be used as the basis for updated mill process design. These were appended to the 2006 Feasibility Study.

Klohn Crippen has completed the detailed design of a compacted cyclone sand tailings dam for the Project, which is expected to be sufficient to support the planned operations over the duration of mine life. It will be sited downstream of the proposed mill site location. Waste dumps have been sited so as to be compatible with plans for surface water management, including a seepage


recovery dam and pond that is located downstream of the main dam structure. Testing of tailings supernatant water shows that most elements meet the British Columbia Water Quality Guidelines (BCWQG) for the protection of freshwater aquatic life.

The 2006 Feasibility Study was based on a 20,000 tonnes per day conventional mill process, fed with ore from an open pit mine. While this overall design concept has not changed in the 2007 Feasibility Study Update, the on-going detailed engineering is considering an average milling rate of 1100 tonnes per hour, as opposed to the previous rate of 906 tonnes per hour.

To take advantage of the expected higher metal prices in the early years of the Project, an initial phase of the mining operation has been included in the mining schedule which will maximize molybdenum production during the first four years of operation. During this phase of the mining, the mine cut-off grade was raised to 0.06 %Mo. However, material grading between 0.04 and 0.06 %Mo would be stockpiled for processing at a later date. By processing a higher grade feed to the mill during this phase, net revenues are maximized allowing a faster pay-back of the initial capital investment.

The total initial capital cost for the development of the Project is estimated to be CDN$640 million. The cost estimate has been carried out to an accuracy of +15%. The estimate was updated from the 2006 Feasibility Study.

A summary of the major capital costs is shown in Table 1.3.

Table 1.3 Capital Cost Summary

Description Estimated Cost (CDN$)

Project Development and Infrastructure 39,700,000 Facilities Construction and Commissioning 341,300,000 Materials and Equipment 134,500,000 Engineering and Project Management 48,800,000 Subtotal 564,300,000 Contingency 55,600,000 Construction Risk 20,100,000 Total 640,000,000

Mining costs were developed by Golder and Adanac from first principles and with various equipment suppliers to determine the most appropriate operating costs for the mining operations. Estimated hourly equipment operating costs developed were compared to actual costs at similar


mines. Labour costs were estimated from information collected from other similar mining operations.

The mine has been designed and costed as an owner-operated mine. The average unit mining cost was determined to be CDN$1.39 per tonne mined or CDN$3.94 per tonne of ore milled, for the first five years of full production and CDN$2.47 for the remaining years of operation.

Process operating supply costs are based on budgetary prices from vendors of the consumables and reagents.

The costs for General and Administration (G&A) includes mine management, transport, insurance, warehouse and security personnel and general management.

Tailings dam maintenance costs have been estimated from published costs from other mines using similar construction techniques.

The total operating cost for mining is estimated to be CDN$13.08 per tonne of ore milled for the first five years of full production and CDN$8.11 per tonne of ore milled for the remaining years of operation. Commissioning and pre-stripping activities completed in Year 1 have been accounted for in the capital cost estimates. The operating costs for mining, processing, power, tailing dam operation, general administration and Adanac’s related cost were prepared by in house professional engineers and reviewed by independent qualified persons to confirm that the work conforms to good engineering practice. Tables 1.4 and 1.5 present the operating cost summary for first four years and after five years of full production to end of life (year 21).

Table 1.4 Operating Cost Summary

(First five years of full production)

Description Operating Cost (CDN$/tonne of ore milled)

Mining (average) 3.94 Processing 2.40 Power 5.57 Tailings Dam Operation (average) 0.26 G&A 0.10 Owner Cost 0.81 Total Average Operating Cost: 13.08



(After Five Years of Full Production)

Description Operating Cost (CDN$/ tonne of ore milled)


CPM Group prepared a detailed market study and price forecast for molybdenum in US$/lb in their October 2007 CPM Marketing Report. This is presented in Table 1.6.

Table 1.6 Molybdenum Price Outlook

Year Price per pound (US$)

2008 34.00 2009 32.25 2010 28.00 2011 23.00 2012 21.75 2013 19.50 2014 16.00 2015 15.00

2016 onwards 14.75

Estimates of production, operating costs, projected metal prices and possible gross margins for the first four years of production are summarized in Table 1.7.


Table 1.7 Production and Operating Costs (First Four Years of Production)

Year lbs Mo x 106 Total

Cost/Tonne (CDN$)

Total Cost/lb Mo

(US$)

Projected Price

(US$/lb)

Projected Gross Margins (US$/lb)

2009 8.143 13.48 7.42 32.25 24.83

2010 13.816 12.25 6.29 28.00 21.71

2011 12.463 12.38 6.87 23.00 16.13

2012 12.147 13.68 7.78 21.75 13.97

After the initial four years of production, the mine cut-off grade is predicted to reduce to 0.04 %Mo and the operating plan will focus on maximizing overall project cashflow.

Adanac's goal is to become a producer as soon as possible as the timing of the project is critical to fully take advantage of the current high molybdenum prices. The decision has been made to initially use diesel-electric power generation in order to expedite the proposed development schedule. This has added significantly to both capital and operating costs but enables the project to start-up at least two years earlier than otherwise could be achieved. Adanac believes that connection to grid electric power from Yukon will occur by the end of 2013.

Salient points of the overall study are:

• Mine Life: 21 years

• Milling Rate: 23,000 tonnes per day

• Strip Ratio: 1.11 (tonne waste)/1.0 (tonne ore)

• Tonnage Milled: 157.6 million tonnes, average grade 0.058%Mo

• Molybdenum in concentrate: 81.7 million kilograms

• Pre-production Capital: CDN$640.0 million

• Average Operating Cost: US$7.60/lb Mo for first five years of full production and. US$7.99/lb Mo afterwards

A pre-tax economic model has been developed from the estimated costs and the open pit production schedule. The Base Case has an Internal Rate of Return (IRR) of 18.9% and a Net Present value (NPV) of CDN$295.0 million, at an 8% discount rate, with a 21-year mine life. The payback of the initial capital is 3.2 years.


The timing of the Project is critical if Adanac is to take advantage of the higher molybdenum prices early in the scheduled mine life. Additional economic sensitivities were run for variations in price, capital cost and operating cost, and the comparative indicators are summarized in Table 1.8.

Table 1.8 IRR and NPV Summary Results

Case Description IRR NPV @ 8%

(CDN$ millions)

Payback Period (Years)

Base Case 18.9% 295.0 3.2 Sensitivities Historical Average Mo Price (Last 3 years’) 30.3% 1,014.7 2.9 Low Case Mo Price Scenario 12.3% 120.7 5.9 High Case Mo Price Scenario 24.6% 444.1 2.6 Capital Cost +15% 14.8% 213.0 3.8 Capital Cost -15% 24.8% 377.2 2.6 Operating Cost +15% 15.2% 185.6 3.5 Operating Cost -15% 22.4% 404.5 2.9 Specific Economic Sensitivities Mine operation with hydroelectric power starting earlier in year 3 (Base Case is year 6) 19.5% 325.0 2.6

Increase in in-situ grade by 15% 27.6% 533.4 2.4

Detailed engineering and procurement have been underway since November 2006. Adanac announced a positive production decision subsequent to the receipt of the Environmental Assessment Certificate on September 10, 2007 from the Province of British Columbia. Adanac is in the process of arranging equity and debt financing to build the mine.

Based on the results of this analysis, the Project contains a valuable molybdenum-bearing mineral resource that can be economically extracted using proven mining methods and processing technologies, at current labour, equipment and material costs and also based on the projected prices of Molybdenum in the future.

It is recommended that Adanac continue to develop the Project through detailed engineering and construction.


2.0 INTRODUCTION

Adanac Molybdenum Corporation (“Adanac”) is currently evaluating the development of the Ruby Creek Molybdenum Project (“the Project”). The Project is planned as an open pit mining operation with on-site ore beneficiation. The mineral resource is a porphyry molybdenum deposit in multiple phases of felsic intrusions. Based on the ore characteristics and reserve magnitude, the mill has been designed to operate at an average rate of approximately 23,000 metric tonnes per day for an life of at least 21 years. The site has no developed infrastructure with the exception of a single lane access road.

Mining and processing are scheduled on a 24 hour per day, 7 days per week schedule. The project will be operated as a fly in/fly out camp accommodation program, with crews working on a two week turn-around basis. The operating management and labour positions are expected to peak at 250 employees.

This report was prepared by Rick Alexander (“Alexander”) at the request of Adanac, in order to prepare a feasibility study to incorporate all new information on the project including current resource and reserve estimates and to outline the standards of disclosure required for mineral projects under National Instrument (“NI”) 43-101. The work has included a Mineral Resource estimate, mine design and the development of a metallurgical process. This NI 43-101 compliant Technical Report has been based on work by Adanac’s in-house professional staff, G&T, SGS MinnovEX, Wardrop, Golder, Klohn Crippen, and CPM Group.

The report updates and summarizes the technical content from several previous reports, as follows:

• Golder Associates Ltd., Ruby Creek Molybdenum Project Mining Feasibility Study, March, 2006 (Qualified Person: Kirk Rodgers, P.Eng.).

• Golder Associates Ltd., 2007 Mineral Resource Estimate Update Ruby Creek Molybdenum Project in British Columbia, Canada, July 23, 2007 (Qualified Person - Paul Palmer, P.Eng., P.Geo.).

• Golder Associates Ltd., Pit Slope Stability Considerations for the Ruby Creek Project, Adanac Moly Corp, Atlin, BC, February 08, 2007 (Qualified Person - Al Chance, P.Eng.).

• Wardrop Engineering Inc., Ruby Creek Feasibility, Process and Infrastructure Design and Cost Estimate, March 2006 (Qualified Person - Rick Alexander, P.Eng.).

• Klohn Crippen Consultants Ltd., Ruby Creek Project — Feasibility Design of Tailings Facility, Waste Dumps and Site Water Management, February 8, 2006 (Qualified Person-Howard D. Plewes, P.Eng.).


• Klohn Crippen Consultants Ltd., Ruby Creek Project — Site Water Management Design Report, February19, 2007 (Qualified Person-Howard D. Plewes, P.Eng.).

• Tailings Facility Detailed Design Report, Kohn Crippen Berger Ltd., April 5, 2007 (Qualified Person-Howard D. Plewes).

Golder has also performed additional technical work in the area of mineral resource estimation.

Alexander the Qualified Person within the Meaning of NI 43-101 has prepared, or supervised the technical matters covered by this report.

A site visit inspection was completed by Alexander on June 21, 2005 for one day and on several occasions between February and December 2007, site visits by Golder qualified person’s have been completed in 2005, 2006 and 2007 to review the data collection procedures and a review the results from the sampling quality assurance and quality control (QA/QC) program for molybdenum and trace element assaying that is completed by Adanac.

Currencies are expressed in Canadian Dollars unless identified otherwise. (United States Dollars: US$.


3.0 RELIANCE ON OTHER EXPERTS

I have relied upon, and believe that I have a reasonable basis to rely upon, the marketing information contained in the molybdenum market study and economic model prepared for Adanac by CPM Group.

With respect to the ownership of the mineral and placer tenures described in Section 4.0 – Property Description and Location, I have relied upon, and believes that I have a reasonable basis to rely upon, the title opinion of Fraser and Company LLP, (November 30, 2007).


4.0 PROPERTY DESCRIPTION AND LOCATION

The Ruby Creek property is in the Atlin Mining Division and originally consisted of a single, irregularly shaped block of 20 claims, comprising 189 units, covering the upper southwest part of the Ruby Creek valley and much of the adjacent Boulder Creek valley. The claims (numbered 510307, 510308, 510309, 510310, 510315 and 530317) were converted into a Mining Lease on March 27, 2007 to facilitate project development.

Subsequently, Adanac has acquired selected mineral and placer tenures adjacent to the Mining Lease or in it’s vicinity, to compliment its original holdings. All mineral tenures are 100% owned by Adanac and are not subject to any royalties or carried interests.

The property is located 24 km northeast of Atlin, British Columbia. Atlin is 175 km southeast of Whitehorse, Yukon Territory. The approximate geographic centre of the original claims on the National Topographic 1:50,000 Map Sheet 104N/11 is universal transverse mercator (“UTM”) Zone 8, 590,000 m east and 6,620,000 m north. Figure 4-1 shows the location map of the project. Six claims were converted into a Mining Lease on March 27, 2007.

Figure 4-1 Location Map


Fraser and Company LLP (“Fraser”) is of the opinion that the recorded owner and expiry periods for the mineral and placer claims are as listed in Table 4.1.

Table 4.1 List of Mineral Claims


* 140154 is the owner number of the Corporation, who is the recorded owner of these claims. Fraser’s search does not indicate any recorded liens, encumbrances or agreements against the above claims. The records of the MTO (which is the source of this information) are subject to the provisions of the Mineral Tenure Act (British Columbia) and the regulations thereunder. Fraser confirms that a mining lease over the following six mineral tenures has been issued as of March 27, 2007 and granted for an initial term of 30 years. The mining lease number is 555153:

- 510307 - 510308 - 510309 - 510310 - 510315 - 510317

The title can be affected by a number of other factors, which the Mineral Titles Office would have no record of and there is no statutory requirement to record with the Mineral Titles Office bills of sale, agreements or other documents that may affect title to the mineral claims. The original mineral claim package is shown in Figure 4-2. The main mineralization is located on the Mining Lease in the northeast area of these claims. This information was provided by Adanac.


Figure 4-2 Mineral Claims Map1

1Provided by Adanac

Approximate Ruby Creek

Property Boundary

N


5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

The resource is a molybdenum porphyry deposit located at the headwaters of Ruby Creek, which flows south into Surprise Lake. The property lies at an elevation of about 1,400 metres, and the deposit underlies a relatively unvegetated alpine cirque, near the head of the valley. The walls are moderately steep but the floor is glacially scoured and flat. The deposit is easily accessed by 40 km of road from Atlin. The first 19 km to the bridge at Surprise Lake is fully maintained; the remaining 21 km is a single lane exploration road that is not maintained.

Other than road access, there is no infrastructure on the property other than a core logging building and core storage racks. The town of Atlin, located 24 km southwest of the property, is a source for fuel, groceries, accommodation, charter aircraft services, etc. Atlin is accessible by a two-hour drive on an all-weather road south from Whitehorse, Yukon, the territorial capital and major supply centre for the region. Whitehorse is serviced by daily commercial flights to Vancouver and other cities.

The Ruby Creek area lies east of the Coast Range Mountains and within a zone generally described as having an interior type of climate. In general, the winters are severe and the summer months are cool. Summer is enhanced by long hours of daylight; during June and July, daylight lasts up to 20 hours.

Temperature was measured concurrently at Atlin and Ruby Creek in 1979 and 1980. A weather station was installed on the property (June 2004) to provide meteorological information needed for environmental baseline studies.

Atlin long-term temperature records indicate:

• Mean annual temperature is about 1°C;

• Mean daily temperatures are above freezing from April to October;

• Temperatures range from -50°C to +31°C; and

• From the limited comparative weather information (Atlin and Ruby Creek) the site has:

- Mean annual temperature that is approximately -2°C, which is 3°C lower than Atlin; - Mean summer temperatures that are about 5°C lower than Atlin; - Mean daily temperatures (June to September) are usually above freezing; and - Temperature extremes that range from -53°C to +28°C.

Freezing temperatures can be encountered at any time of the year in Atlin and Ruby Creek.


Precipitation data at the Project site is available for the summer months of 1979 and 1980, and from the weather station installed in June 2004. Comparison of the summer rainfall at Atlin and Ruby Creek during 1979 and 1980 indicated the Ruby Creek had 1.6 to 2.2 times the rainfall measured at Atlin. The estimated mean annual precipitation at Ruby Creek is 715 mm.

The surface rights cover an approximate area of 24.65 square kilometres and therefore is of sufficient size to allow mining operations to be carried out.


6.0 HISTORY

The history of the Ruby Creek deposit has been provided by Pinsent (2005) and previous Technical Reports (Blower, 2005; and Palmer, 2006 and 2007). Excerpts from those reports are included below.

The Ruby Creek deposit has been known since 1905, but underwent only limited exploration until 1966, when it was staked by Adanac Mining and Exploration Limited (no relation to Adanac) and Canadian Johns Manville Limited. Adanac Mining and Exploration Limited acquired the controlling interest the following year and cored an aggregate length of 12,775 m in 80 drill holes. It optioned the property to Kerr Addison Mines Limited (Kerr Addison), in 1970.

Kerr Addison completed a further 47 diamond drill holes for a total length of 5,626 m and drove 589 m of drift, 246 m of cross-cut and 281 m of raise in the high-grade core of the deposit. It extracted 9,545 tonnes of mineralization, primarily from six raises and processed it on site to assess the significance of a well-defined nugget effect caused by coarse-grained molybdenite mineralization and confirm a viable metallurgical process. Chapman, Wood and Griswold Limited completed a feasibility in 1971 and deemed the deposit un-economic on account of prevailing low prices for molybdenum (US$1.90/lb). As a result, Kerr Addison dropped its option.

The following year, 1973, Climax Molybdenum Corporation of British Columbia Limited (Climax) drilled and/or deepened a further 9 diamond drill holes for an aggregate length of 2,672 m and developed the first comprehensive geological model for the deposit (White et al., 1976). Climax dropped its option and thereafter, the property remained dormant until metal prices improved in the late 1970s.

In 1978, Placer Development Limited (“Placer”) optioned the property and initiated a full-scale technical and socioeconomic evaluation. In 1979, they completed a further 6,028 m of diamond drilling in 49 holes in-and-around Kerr Addison's proposed initial pit area, and the following year drilled a further 27 diamond drill holes with a total length of 4,858 m, in the ultimate pit area. Although Placer carried out nearly all of the work required for a feasibility study, the price of molybdenum dropped sharply in 1983 and the project was shelved. The company held on to the option for a few years but eventually returned it to Adanac Mining and Exploration Limited.

The claims lapsed in the late 1990s and Andris Kikauka staked the deposit for Adanac Gold Corp, now Adanac Molydenum Corp. (“Adanac”) in 2002. The following year, Adanac started to compile the considerable amount of historic data that documented the previous work on the property. Following a positive scoping study early in 2004, Adanac did further drilling on the property. The objectives were to fill gaps in existing data distribution, to get new assay


information in order to assess the quality of the old assay data and to improve the understanding of the deposit. The programme was designed with input from AMEC Americas Ltd. (“AMEC”) and, later in the same year, Adanac completed 38 diamond drill holes totalling 9,087 meters.

In April 2005, the mineral resources for the Ruby Creek deposit were outlined in the Technical Report – Mineral Resource Estimate Ruby Creek Project (Blower, 2005) by AMEC.

During 2005, Adanac completed a 19 diamond drill hole program, totalling 4,984 m, based on recommendations in the Technical Report – Mineral Resource Estimate Ruby Creek Project (Blower, 2005) and with additional input from Golder. Of the 19 diamond drill holes, 7 were drilled within the proposed pit area to infill past campaigns, obtain metallurgical samples and 12 [including five dual purpose holes (exploration / geotechnical)] were drilled on the extremities of the property to delineate the fringes of the deposit and to provide geotechnical data for the proposed pit slope design. The geotechnical holes included two holes drilled in the south wall, one hole in the west wall, one hole in the northwest wall, and one hole in the north to intersect the Adera Fault area

During 2006, Adanac completed a 16 diamond drill hole program totalling 3,921 m in and around the previously established deposit (central deposit area) as well as the south and southwest edges of the deposit. Three of the holes were located on the fringes of the deposit to improve Adanac’s understanding of the overall shape, depth and extremities of the deposit. The remaining 13 holes were located to infill past campaigns and drilled at approximately -50° dip and 270° Azimuth. In addition, during the 2006 drilling program, 8 drill holes (AD-357, AD 359, AD-361, AD-363, AD-364, AD-366 to AD-368) from the inclined drilling program were surveyed with an optical televiewer camera to determine dominant dip and dip directions of the mineralized veins.

The 2006 database for the Ruby Creek deposit is composed of 266 drill holes with a total of 46,912 m of drilling information. Between the years of 2004 and 2006, Adanac has drilled a total of 73 of the 266 drill holes.

In 2006 a NI 43-101 Technical Report was completed based on the 2005 data and earlier and was entitled “Mineral Resource Estimate Ruby Creek Molybdenum Project” and dated February 8, 2006 (“February 8, 2006 Mineral Resource Estimate”).

In late 2006 and early 2007 the 2006 Datamine Database was used in developing the updated Mineral Resource Estimate for the property and is entitled “2007 Mineral Resource Estimate Update Ruby Creek Molybdenum Project in British Columbia, Canada”, dated July 23, 2007 (July 23, 2007 Mineral Resource Update”). The July 23, 2007 Mineral Resource Update was used in the development of this Feasibility Report which is based on 266 drill holes (71 from Adanac) drilled between 1966 and 2006.


Outlined in Table 6.1 is the previous NI-43-101 2006 Mineral Resource Estimate for Ruby Creek based on the 2006 Technical Report (Palmer, 2006).


Table 6.1 January 2006 Mineral Resource Estimate

Ruby Creek Molybdenum Project

Resource Category Cut-off (%Mo) Tonnage %Mo Mo lb 0.020 40,636,000 0.077 68,982,000

0.030 40,386,000 0.077 68,557,000

0.040 38,942,000 0.079 67,822,000 0.050 35,834,000 0.081 63,990,000

0.060 28,836,000 0.088 55,942,000

0.070 22,593,000 0.094 46,820,000

0.080 15,225,000 0.104 34,907,000

0.090 10,601,000 0.112 26,175,000

Measured

0.100 6,908,000 0.121 18,427,000

0.020 432,936,000 0.041 391,325,000

0.030 261,618,000 0.051 294,149,000

0.040 167,433,000 0.059 217,782,000 0.050 113,435,000 0.066 165,052,000

0.060 63,101,000 0.076 105,725,000

0.070 37,572,000 0.083 68,751,000

0.080 17,543,000 0.094 36,355,000

0.090 8,808,000 0.103 20,001,000

Indicated

0.100 4,151,000 0.113 10,340,000

0.020 473,572,000 0.044 460,307,000

0.030 302,004,000 0.054 362,706,000

0.040 206,375,000 0.063 285,604,000 0.050 149,269,000 0.070 229,042,000

0.060 91,937,000 0.080 161,667,000

0.070 60,165,000 0.087 115,571,000

0.080 32,768,000 0.099 71,262,000

0.090 19,409,000 0.108 46,176,000

Measured + Indicated

0.100 11,059,000 0.118 28,767,000

0.020 151,326,000 0.034 113,429,000

0.030 61,837,000 0.048 65,437,000

0.040 33,067,000 0.060 43,740,000 0.050 23,225,000 0.067 34,305,000

0.060 13,375,000 0.076 22,409,000

0.070 6,490,000 0.088 12,591,000

0.080 3,166,000 0.102 7,120,000

0.090 1,915,000 0.113 4,771,00

Inferred

0.100 1,143,000 0.124 3,124,000 Cut-off %Mo grades were classified as greater than or equal to and range from 0.01 to 0.10 in increments of 0.01. The mineral resource estimate has been completed in accordance with CIM standards of Estimation of Mineral Resources and Mineral Reserves.


7.0 GEOLOGICAL SETTING

The regional and local geological setting descriptions of the Ruby Creek deposit are provided in reports by Pinsent (2005), Blower (2005) and Golder’s Mineral Resource Estimate Reports, (2006 and 2007). Excerpts from those reports are provided below.

In Golder’s opinion, the geology of the Ruby Creek deposit and the mineralization controls are sufficiently well understood and sufficiently reliable to be used in the development of the resource estimation.

Additional analysis has been completed during the 2006 drilling program in order to better understand the sub-vertical and sub-horizontal vein orientations and molybdenum mineralization using in-hole optical televiewer cameras. The 2006 drilling results indicated that high grade mineralization did occur along both sub-vertical and sub-horizontal veins. These vein orientation results were also applied to the mineral resource estimate process. However, more inclined drilling data is required to better understand the relationship between sub-vertical and sub-horizontal vein orientations and higher grade molybdenum mineralization.

7.1 Regional Scale

The Ruby Creek deposit is a disrupted, dome-shaped occurrence formed late in the development of a localized plutonic complex. It is associated with granitic to quartz monzonitic rocks of the Surprise Lake Batholith, east of Atlin. Regional geology is shown in Figure 7-1.


Figure 7-1 Regional Geology Map of the Atlin Area

The geology of the Atlin area was mapped by Aitken (1959) and the regional setting of the deposit is discussed by Christopher and Pinsent (1982). Described simply, the Atlin area is underlain by deformed and weakly metamorphosed ophiolitic rocks of the Pennsylvanian and/or Permian-aged Cache Creek Group (Monger, 1975). These rocks, which include serpentinites and basalts as well as limestones, cherts and shales, are thought to be the source of much of the placer gold found in the Atlin area. The stratigraphic rocks are cut by two younger batholiths. North of Pine Creek, the stratigraphy is cut by a Jurassic-age granodiorite to diorite intrusion known as the Fourth of July Batholith and, north and south of Surprise Lake, it is cut by a Cretaceous-age granitic to quartz monzonite intrusion known as the Surprise Lake Batholith.

The rocks are locally strongly faulted and the Ruby Creek deposit is located near the intersection of two major, pre- to post-mineral fault systems. The deposit location is partially offset by the Adera fault system which trends from southwest to northeast down Ruby Creek and defines much of the southern boundary of the fourth of July Batholith. The deposit is also controlled by the Boulder Creek fault system. This runs due north up Boulder Creek and cuts across the head of the Ruby Creek drainage. The Boulder Creek fault appears to have helped localize emplacement of the deposit, which is intimately associated with late-stage porphyritic and aplitic plutonic rocks intruded into a marginal phase of the Surprise Lake Batholith.


Ruby Mountain, immediately to the south of the deposit, is underlain by Late Tertiary to Quaternary flows from a volcano that erupted and filled the lower part of the Ruby Creek drainage with columnar basalt and volcanoclastic debris. The volcanic rocks unconformably overlie placer gold-bearing gravels. The origin of the gold is uncertain; however, most of it probably comes from quartz-carbonate veins hosted by shears that cut Cache Creek Group strata.

7.2 Local and Property Scale

The Ruby Creek deposit underlies the valley floor near the head of Ruby Creek. It is largely buried and has very little surface expression. There is little outcrop in the lower part of the valley and molybdenite is only rarely found in float and/or in veins outcropping in the bed of the creek. The geology underlying the valley floor is largely derived from drill data.

The Ruby Creek area is underlain by two separate pulses of plutonic rock. The first pulse, which includes the contact phase between the two batholiths, consists of a highly variably textured unit that grades from Coarse-grained Quartz Monzonite (CGQM) south of the Adera fault through a number of texturally transitional phases including Transitional and/or Hybrid Coarse-grained Quartz Monzonite (CGQM-T; CGQM-H) and Crowded Quartz Feldspar Porphyry (CQFP) to Sparse Quartz Feldspar Porphyry (SQFP) upward and outward from the deposit. The latter is well exposed north of the Adera fault, near the diorite contact.

The CGQM is weakly to moderately deformed pink to grey equigranular, coarse grained (0.5 to 3.0 cm) quartz monzonite consisting of equal amounts of orthoclase, plagioclase and grey quartz (Christopher and Pinsent, 1982). The feldspar is commonly seriate and, locally, includes a small amount of fine-grained (2 to 4 mm) matrix. CGQM grades to SQFP with increase in matrix content, and increased isolation of constituent phenocrystic crystals, particularly orthoclase and quartz.

The first phase also includes a distinctive Mafic Quartz Monzonite Porphyry (MQMP) unit that is present east of the deposit. This distinctive grey rock type has a seriate (1 to 4 mm locally) porphyritic texture. It is composed largely of chalky white plagioclase, disseminated biotite and phenocrysts of ragged plagioclase and lesser quartz. These rocks were fractured and deformed prior to emplacement of the second pulse of magma.

There are three main mappable phases to the second pulse. They include Crowded Quartz Monzonite Porphyry (CQMP), Sparse Quartz Monzonite Porphyry (SQMP) and Fine-grained Quartz Monzonite (FGQM).

The CQMP has an average of 50% (2 to 6 mm) subhedral to euhedral plagioclase, orthoclase, quartz and biotite phenocrysts in an aphanitic matrix. The SQMP variety is similar, but has fewer


(10% to 30%) phenocrysts. The SQMP is fresher and generally less deformed than the surrounding rocks and has a much finer, more chilled matrix than the sparse quartz feldspar porphyry described above. The second phase porphyries cut out the older rock units and are exposed locally in the floor of the valley. They are also found in the subsurface, under the valley floor, upstream where the CGQM and its variants are intruded by a buried cupola of SQMP. Its shape has strongly influenced the locus of mineralization, as shown by Placer's 0.06% Mo and 0.1% Mo assay contours at 1448 m elevation. Mineralization surrounds the buried cupola and, to a lesser extent, covers it.

The FGQM is a variably textured aplite that intrudes the CGQM (and also its variants) and the MQMP, above and around the sparse and crowded porphyry intrusions. This rock type is not exposed on the surface, but it is well documented as forming a series of 0.05 to 10 m thick, approximately flat lying, structurally-controlled sills in the higher-grade (north-eastern) portion of the deposit. FGQM dykes are found elsewhere around the buried sparse-porphyry cupola; however, they are generally less frequent and smaller, and occur as narrow dykelets.

In addition to these rock types, recent drilling at the southwest end of the deposit has located a Megacrystic Porphyry (MFP) unit in the subsurface. This is not well constrained; however, it appears to be a relatively young phase of the quartz monzonite intrusion. It consists of rare to abundant large (> 10 mm) euhedral orthoclase phenocrysts in a chilled matrix. Another notable feature throughout the deposit is the presence of coarse-grained quartz-feldspar pegmatite. This is not abundant, but covers a wide area as small dykes and structurally controlled sills.

Large-scale fault structures and their splays have provided conduits for mineralizing fluids and have localized mineralization. The deposit is situated at the intersection of the Adera and Boundary Creek faults. The Adera fault is particularly important because it offsets the northern portion of the deposit. It is a composite structure that dips steeply to the northwest, is normal in character and appears to have down-dropped (to the north) the north-western part of what was originally a dome, or ring-shaped deposit formed above and around a sparse-porphyry intrusion. Mineralization has been found in the coarse-grained and related rocks northwest of the southernmost strand of the Adera fault. However, it has not been found in (probably similarly aged) sparse quartz feldspar porphyry and related rocks further north. These rocks, which are well exposed in the creek canyon below Molly Lake, contain abundant barren quartz veins and disseminated pyrite. They are gossanous, but barren.

In addition, during the 2006 drilling program, 8 drill holes (AD-357, AD 359, AD-361, AD-363, AD-364, AD-366 to AD-368) from the inclined drilling program were surveyed with an optical televiewer camera to determine dominant dip and dip directions of the mineralized veins. This work was completed by the Golder Burnaby office as a separate project from the mineral resource estimate. At the time of the July 23, 2007 Mineral Resource Update, a review of this data


collected by Golder was reviewed by the qualified person and three dominant vein orientation sets were identified and are as follows:

• Set 1: 50-80° dip/170-195° dip direction;

• Set 2: 50-85° dip/300-310° dip direction; and

• Set 3: 5-15° dip/350-010° dip direction.

The structural data collected from the inclined holes were based on veins sizes ranging from <5 mm, 5-10 mm and >10 mm. Set 3, near horizontal veins, had the largest number of veins identified in the analysis, which is consistent with the current understanding of mineralization. Sub-vertical mineralization veins (Sets 1 and 2) were identified during the structural analysis, but previously of unknown orientation.

It is likely that Set 2 could be a subset of the Adera Fault since it is estimated as dipping steeply to the northwest.

The rocks at Ruby Creek are, for the most part, fresh and much of the alteration that is observed is post-mineralization, associated with fluids that circulated during post-mineral faulting. However, there is a small amount of primary alteration. It occurs as sill-like zones of intense silicification intermixed with bodies of aplite in the higher-grade, north-eastern part of the deposit and as intermittent feldspar envelopes and flooding around individual mineralized quartz veins throughout the deposit. In one locality, the silicification can be shown to pre-date both intrusion of aplite and emplacement of mineralized quartz veins.

Fractured rocks near post-mineral faults, such as the Adera, have commonly undergone late hydrothermal alteration. They are either weakly or strongly altered to a mixture of sericite, carbonate, clay and chlorite (without addition of secondary quartz). The altered rocks become soft and friable and early (1969/72) core recoveries were lower in these areas. Major faults commonly include breccias cemented by grey gouge of similar composition, with or without smeared molybdenite. Some of the altered rocks contain fluorite veins. Work by Placer in 1980 shows that most of the light-coloured clay is predominantly montmorillonite; however, the grey clay in the main Adera fault zone consists largely of kaolinite.


8.0 DEPOSIT TYPES

Details of the deposit type of the Ruby Creek property are contained in Golder’s Mineral Resource Estimate Reports, (2006 and 2007) which are based on the reports by Sinclair (1995) and Blower (2005). Excerpts from those reports are provided below.

Ruby Creek can be classified as a low fluorine porphyry molybdenum deposit (Sinclair, 1995). These deposits are characterized by stockworks of molybdenitebearing quartz veinlets and fractures in intermediate to felsic intrusive rocks and associated country rocks. They are typically low-grade but large and amenable to bulk mining methods.

Porphyry molybdenum deposits vary in shape from an inverted cup, to roughly cylindrical, to highly irregular. They are typically hundreds of metres across and range from tens to hundreds of metres in vertical extent. Mineralization is predominantly structurally controlled, consisting mainly of stockworks or crosscutting fractures and quartz veinlets, with veins, vein sets and breccias. Molybdenite is the principal ore mineral; chalcopyrite, scheelite, and galena may be present but are generally subordinate (Sinclair, 1995).

These deposits are thought to originate from large volumes of magmatic, highly saline aqueous fluids under pressure. Multiple stages of brecciation related to explosive fluid pressure release from the upper parts of small intrusions result in deposition of ore and gangue minerals in crosscutting fractures, veinlets and breccias in the outer carapace of the intrusions and in associated country rocks. Incursion of meteoric water during waning stages of the magmatic-hydrothermal system may result in late alteration of the host rocks, but does not play a significant role in the ore-forming process (Sinclair, 1995).


9.0 MINERALIZATION

A detailed description of the mineralization is contained in Golder’s Mineral Resource Estimate Reports (2006 and 2007), which is a summary based on reports by Pinsent (2005) and Blower (2005). Excerpts from those reports are provided below.

The Ruby Creek deposit consists of a stockwork of veins of molybdenite and quartz molybdenite found in all the principal rock types. However, it is best developed in the early stage plutonic rocks (mafic quartz monzonite porphyry and coarse grained quartz monzonite and its variants) that overlie and surround the buried sparse quartz monzonite porphyry stock under the Ruby Creek valley. The veins are most commonly without other metallic phases, although pyrite is found locally and chalcopyrite has been observed. The veins locally contain traces of scheelite, orthoclase, fluorite, biotite, sericite, and carbonate.

Mineralization post-dates emplacement of fine grained quartz monzonite in the higher-grade zone, located on the northeast side of the deposit. In this area, there appears to be a crude positive correlation between the presence of dykes and sills and the amount of mineralization observed. However, the same relationship does not hold on top of the cupola or in the south-western part of the deposit. The deposit consists of a mineralized blanket that covers the sparse quartz monzonite porphyry stock and dips off in all directions.

The mineralization commonly consists of sulphide veins as coatings on quartz free fractures, and as coarse and fine rosettes and blebs in both smoky and lesser clear quartz. It also occurs as streaks and smears in deformed rock and may, locally, be enriched in fault zones. In the higher grade zone, explored by Kerr Addison, much of the mineralization is in horizontal to sub-horizontal veins and fractures from 1 mm to 5 mm wide that are interspersed with veins that are considerably wider, up to 20 mm wide. The narrow quartz veins are oriented at a high angle to the (vertical) core axis. Both vein sets are mineralized and blebs of molybdenite commonly occur at the intersection of cross cutting veinlets.

The near horizontal vein set is locally extremely well mineralized. Veins exposed through underground development in the 1970's show that coarse rosettes of molybdenite up to 30 mm in diameter formed in the plane of the vein, and that the spacing between the rosettes is variable, causing a pronounced nugget effect in drilling.

The crowded and sparse porphyries underlying the higher grade zone are cut by narrow (1 mm to 3 mm) mineralized quartz veins and fractures that also occur at both high and low angles to the (vertical) core axis. These veins and fractures commonly contain fine grained to powdery molybdenite. There are fewer high grade rosettes formed at depth.


As outlined previously, optical televiewer surveys were completed in eight inclined boreholes form the 2006 drilling program in order to determine the dominant orientation of veins in the central portion of the deposit. The following three dominant sets have been identified from the survey:

• Set 1: 50-80° dip/170-195° dip direction;

• Set 2: 50-85° dip/300-310° dip direction; and

• Set 3: 5-15° dip/350-010° dip direction.

The dominant vein set identified from the survey is the horizontal set followed by two sub-vertical sets, Set 1 dipping south and Set 2 dipping northwest. It was identified by the Adanac geology staff that zones that were dominated by both horizontal and vertical veins sets in the 2006 drilling program showed an increased molybdenite mineralization. Therefore, the drilling of inclined boreholes in areas where both horizontal and vertical veining occurs increases the chance of sampling these zones. Based on the orientation of sub-vertical sets any future planned inclined boreholes should be drilled perpendicular to dip direction of these sets (azimuth directions north and southeast).


10.0 EXPLORATION

Much of the past efforts (post 1966) at Ruby Creek have focused on the central portion of the deposit. The Placer and Kerr Addison exploration campaigns both focused on the central portion of the deposit to depths approximately 200 m below ground level. Mineralization occurs below this depth, but is not necessarily amenable to open pit mining. Descriptions of the past exploration programs prior to Adanac exploration programs has been described in Section 6 (History). The following sections focus on the exploration programs completed by Adanac from 2004 to 2006

10.1 ADANAC 2004

Adanac conducted a major exploration drilling campaign in 2004. The company drilled 38 holes having an aggregate length of 9,087 m in and around a previously established deposit and submitted 2,830 samples for molybdenum assaying. Additionally, 256 samples were collected for specific gravity testing by ALS Chemex Laboratories (ALS Chemex) in Vancouver, BC.

Ten holes were twinned in 2004 (five twinned with Kerr Addison and five with Placer). These holes were designed to confirm drill holes from earlier campaigns to validate original drilling results. The remaining holes were located to improve Adanac’s understanding of the overall shape of the deposit. The results were reviewed, validated and Mineral Resources in compliance with NI 43-101 standards were first reported in AMEC's Technical Report (Blower, 2005).

10.2 ADANAC 2005

Adanac expanded on their 2004 campaign during 2005. Adanac drilled 19 holes having an aggregate length of 4,984 m in and around the previously established deposit as well as on the deposit fringes. Five of the holes were geotechnical holes designed to generate information necessary for pit slope stability assessment. The remaining twelve holes were located to infill past campaigns on the fringes to improve Adanac’s understanding of the overall shape, depth, and extremities of the deposit.

Drill core sampling for the 2005 drilling program was based on the 2004 QA/QC program developed by Adanac. A total of 1,559 drill core samples were submitted to ACME Analytical Laboratories (ACME) in Vancouver for analysis.

The seven holes internal to the established deposit, AD-337 to AD-343, were submitted for molybdenum analysis. The three short distal exploration holes, AD-344 to AD-346 were submitted for molybdenum as a trace element. The remaining drill holes (including geotechnical holes), AD-347 to AD-355 were also analysed for molybdenum as a trace element. Additionally,


60 samples from seven drill holes (AD-337 to AD-343) were submitted for assay checks to ALS Chemex as part of the quality control program.

A total of 615 samples were also collected for specific gravity testing and were submitted to ALS Chemex. These samples included 332 from 19 drill holes in the 2005 program and 283 from 30 holes in the 2004 drilling program. The results from the specific gravity testing showed similar results to the testing results in the Technical Report – Mineral Resource Estimate Ruby Creek Project (Blower, 2005).

All 2005 drill hole collars were surveyed at the end of the drilling program by Underhill Geomatics Ltd and were provided by Adanac using the NAD 27 UTM co-ordinate system which was also used in the 2004 drilling program. The NAD 27 UTM same co-ordinate system was also used in the historical drilling programs and was used in the 2004 and 2005 drilling programs for consistency. All 5 geotechnical drill holes were inclined dipping and were surveyed by Golder using an optical televiewer system.

The 2005 drilling program, sampling program and surveying was reviewed and validated and, in Golder’s opinion, was sufficient to include in the 2006 Mineral Resource Estimate. Golder also visit the project site during the 2005 drilling program.

10.3 ADANAC 2006

Adanac continued a drill program during 2006. Adanac drilled 16 holes for an aggregate depth of 3,921 m in and around the previously established deposit (central deposit area) as well as the south and southwest edges of the deposit as illustrated on Figure 10-1. Three of the holes were located on the fringes to improve Adanac’s understanding of the overall shape, depth and extremities of the deposit. The remaining 13 holes were located to infill past campaigns and drilled at approximately -50° dip and 270° Azimuth.

The location of the 2006 drill holes is illustrated on Figure 10-1.


Figure 10-1 Plan View of 2006 Drill Hole Collars


Drill core sampling for the 2006 drilling program was based on the 2004 and 2005 QA/QC program developed by Adanac. A total of 1,238 drill core samples were submitted to ACME in Vancouver, BC for analysis, including 295 samples from drill holes AD-356 to AD-368, which were submitted for molybdenum oxide analysis. For quality control purposes, samples from drill holes AD-369 to AD-371 were also analyzed for molybdenum as a trace element and analysed for 40 other elements. Additionally, 186 samples from 13 drill holes (AD-356 to AD-368) were submitted to G&T Metallurgical Services for metallurgical testing. A total of 176 samples were submitted to ALS Chemex for specific gravity testing from the 2006 drilling program. One specific gravity sample was selected approximately every 50 linear ft.

All 2006 drill hole collars were surveyed at the end of the drilling program by Underhill Geomatics Ltd. and provided by Adanac using the NAD 83 UTM co-ordinate system as well as in the NAD 27 UTM co-ordinate system. The NAD 27 UTM co-ordinate system was also used in the historical drilling programs and in the 2004 and 2005 drilling programs for consistency. The survey consulting company also provided 10 historical drill hole locations re-surveyed in the NAD 83 UTM co-ordinate system.

A decision by Adanac to use the NAD 83 UTM co-ordinate system for future mine site construction resulted in the conversion of the entire database to the NAD 83 UTM co-ordinate system. This conversion was based on a combination of drill holes surveyed in NAD 83 UTM and translating the remaining drill hole locations from NAD 27 UTM to NAD 83 UTM by adding 174N (Y), and subtracting 104E (X).

The 13 infill drill holes were inclined and 8 of these were surveyed by Golder using an optical televiewer camera system. Where possible, the drill holes’ steel collar casings were left behind in the drill holes.

Excel spreadsheets of the geological and assaying data for the 2006 drilling program were reviewed by Golder and included in the Ruby Creek Datamine Database. The QA/QC program was reviewed during the site visit by the qualified person and was consistent with the previous site visit and, in Golder’s opinion, was sufficient to include in the July 23, 2007 Mineral Resource Update.

10.4 2007 Drilling Program

At the time of this report the 2007 drilling program was nearing completion. A site visit was completed by a Golder qualified person on September 25, 2007. At the time of the site visit, 5 drill holes had been completed out of a planned total of 14 drill holes (6,839 m). The details (location of drill holes, assay grades, etc.) of this 14 drill holes have not been included in this report nor have any results from this drilling been included in the July 23, 2007 Mineral Resource Update.


11.0 DRILLING

Summarized in Table 11.1 is the history of the drilling programs from 1966 to 2006. The holes were typically diamond drill holes except for a small number of early rotary drill holes. Not all of these drill holes are contained in the Ruby Creek Datamine Database used for the July 23, 2007 Mineral Resource Update.

Table 11.1 Drilling History Of The Ruby Creek Deposit (Dec 31, 2006)

Company Years Drill holes (m)

Adanac Mining and Exploration, & John’s Manville 1966 to 1970 80 12,775

Kerr Addison Mines 1970 to 1972 47 5,626 Climax Molybdenum 1973 9 2,672 Placer Development 1979 to 1980 76 10,886 Adanac Gold (Adanac Moly Corp.) 2004 381 9,0871

Adanac Moly Corp. 2005 192 4,982 Adanac Moly Corp. 2006 16 3,921 Total - 285 49,950

Notes: 1 Includes 2 re-drills of holes 2 Includes 5 geotechnical holes

The majority of the drilling completed to date on the Ruby Creek deposit is vertical dipping and the main mineralization to date has been identified as along sub-horizontal veins except in the central area of the deposit which is a mixture of sub-vertical and sub-horizontal veins. Therefore, in general the mineralized drill intercepts are representative of the true thickness with some exceptions in the areas where mineralization is both sub-vertical and sub-horizontal vein hosted. The exception to this is the inclined drill holes completed in 2005 (geotechnical holes), the inclined vertical drill holes in 2006 to assist in defining vein orientation and the pseudo-drill holes (horizontal) representing the underground sampling that was completed by Kerr Addison.

The drill holes to date have been spaced predominantly along an exploration grid with East-West and North-South grid lines equal to 064° and 154° azimuths using the UTM NAD 27 co-ordinate systems. The section spacing between the grid lines is 100 ft with East 1 representing 100 ft east of baseline West 00. The spacing of drill holes varies, but is typically spaced 300-400 ft (approximately 90-120 m) apart between Sections East 18 to West 28. A higher grade mineralization zone has been identified around the underground development area, central deposit, which has been drilled with a density of approximately 100 ft (approximately 30 m) spacing between Sections East 6 to West 10.


The historical drilling data was originally transferred from paper logs to electronic spreadsheets by Adanac and then entered into a database that was used for the April 2005 Mineral Resource Estimate (Blower, 2005). This database was provided to Golder as ASCII files and was incorporated in the 2006 Ruby Creek Datamine Database. Drill hole information provided by Adanac from the 2004 to 2006 drilling programs, as Microsoft Excel spreadsheets, were incorporated in the Ruby Creek Datamine Database.

The 2006 drilling program added an additional 16 drill holes to this Ruby Creek Deposit as illustrated on Figure 11-1. Figure 11-1 is a plan view of the deposit showing the location of the 266 drill holes from the 2006 Ruby Creek Datamine Database and includes the historical exploration grid and the UTM NAD 83 Grid. In 2006 all drill hole information was converted from the UTM NAD 27 co-ordinate system to UTM NAD 83.

The 2006 Ruby Creek Datamine Database naming convention for drill holes is based on the various drilling campaigns and included four main groups. The original drill hole names have been re-labelled originally during the 2005 Mineral Resource Estimate (Blower, 2005) with the same naming convention continued in the 2006 Ruby Creek Datamine Database. All drill holes with a prefix of KA for Kerr-Addison, CM for Climax Moly, PD for Placer and AD for Adanac (2004, 2005 and 2006 drilling). Table 11.2 summarizes the drill holes compiled in the 2006 Ruby Creek Datamine Database used in the July 23, 2007 Mineral Resource Update.

Table 11.2 Drill Holes in the 2006 Ruby Creek Datamine Database

Campaign Years Drill Holes (m)

Kerr Addison Mines (including Adanac Mining and Exploration, & John’s Manville)

1966 – 1972 105 16,897

Climax Molybdenum 1973 7 1,148

Placer Development Limited 1979 – 1980 66 9,975

Adanac Gold Corporation (Adanac Moly Corp.) 2004 36 8,984

Adanac Moly Corporation 2005 19 4,982

Adanac Moly Corporation 2006 16 3,921

Kerr Addison Mine (underground sampling as pseudo (includes raises) drill holes

1972 17 1,005

Total 266 46,912


The database contains 17 pseudo-drill holes (includes raises) created from the underground sampling program, by Kerr Addison, for a total of 1,005 m. The underground sampling program consisted of representative samples collected from drift rounds in the adits excavated by Kerr Addison. Access to the underground workings is currently not possible. Therefore, a total of 266 drill holes are currently contained within the 2006 Ruby Creek Datamine Database that has been used in the July 23, 2007 Mineral Resource Update.

Site visits by Golder qualified person’s have been completed in 2005, 2006 and 2007, during the active exploration field season at the Ruby Creek Project site. The same drill core logging procedures have been observed during each field season.

All core logging and sampling by the Adanac geologists is first entered on paper logs and later entered electronically onto computers for permanent storage as Excel spreadsheets. After the core was sampled from the core trays, any remaining core (typically half) was stored on wooden-rebar rack structures in their original open core trays. The core rack structures are stored outside, but are protected under wooden roofs. Each core rack structure is labelled with the drill hole name. There is no core stored on site or available prior to the 2004 drilling program. In my opinion, the drilling practice, logging, handling and storage of core employed at the Ruby Creek deposit is standard in the industry and suitable for Mineral Resource Estimation.

Ruby Creek Feasibility Study - 39 - December, 2007

Figure 11-1 Plan View of 2007 Drill Hole Collars


12.0 SAMPLE METHOD AND APPROACH

In the drill programs completed prior to 2004, the standard practice was to crush all core, saving only a small, lithologically representative sample from each 10 ft interval. This approach was taken in order to minimize handling, reduce molybdenite loss through splitting or sawing, and increase the volume of material sampled (hence improving sampling statistics).

No assay samples or representative drill core samples are available prior to the 2004 drilling programs completed by Adanac. All samples that were collected from the 2004 to the present from drilling programs by Adanac have been sawed or split in half and a portion retained to establish an inventory of archived core samples. A few holes were sampled entirely in order to compare the data to historical assay results. These archived split samples are stored on the permanent core racks on the property site.

During the core logging process, samples were selected for assaying. All drill holes were sampled beginning from the bedrock interface. No overburden samples were collected for assaying in the 2004, 2005 and 2006 drill programs. Sample selection was a combination of lithology type and length. The typical sample length was 3.05 m (10 ft) which was the same length as the drill core run and similar to historical sample lengths previously collected. The maximum (4.8 m) and minimum (1.2 m) samples lengths collected in 2004, 2005 and 2006 programs were typically the first or last samples collected in each drill hole.

In practice, the drill core was processed in three ways. The competent sections of the drill holes were either split using a classic hand-cranked core splitter, or sawed using a fast and efficient (Almonte) core saw. Core intervals that were structurally weak and/or too poorly consolidated to split were totally crushed and then passed through a riffle splitter three times to homogenize the sample before being split into to halves. One half was then treated in the same way as the other half-core crushed samples obtained by cutting and/or sawing, and the other was double-bagged and stored as a primary crushed reject.

As part of Adanac’s QA/QC program approximately 5% of the material left in the boxes (split samples) are analyzed as duplicate samples. In addition to the core duplicate samples, several sets of representative samples (0.5 m to 0.1 m long) were collected from the archived split samples for specific gravity, acid generating potential determination and metallurgical testing. A total of approximately 1,000 samples have been collected from the 2004, 2005 and 2006 drilling programs for specific gravity testing (approximately 50 ft spacing). Therefore, the residual core left on site is incomplete. Most of the specific gravity and representative samples taken from the 2006 drill core are in the company office in White Rock, British Columbia.


During the 2004 program, three drill holes were totally crushed to reproduce the sample handling processes that were taken in pre-2004 drilling programs. These holes (AD-303, AD-307, and AD-308) were collared adjacent to pre-existing holes drilled by Kerr Addison (KA-60-1 17) and Placer (PD-221, PD-227).

Although the process of sawing or splitting the core can cause some loss of molybdenite, it is worth noting that the drill hole recoveries returned in the 2004 through 2006 programs were often higher (>95%) when compared to the losses through down-hole erosion that were experienced by either Kerr Addison or Placer.

Golder reported that, the sampling procedures used in the 2006 drilling program are consistent with the 2004 and 2005 practices developed and are consistent with industry practices and were representative of the drill hole data collected.


13.0 SAMPLE PREPARATION, ANALYSES AND SECURITY

13.1 Field Sample Preparation Procedures

13.1.1 Adanac 2004, 2005 and 2006

As discussed in the previous section, all samples selected for assaying were crushed on site to less than approximately 10 mm (3/8") using a Nelson Machinery Atlas core crusher that was cleaned with compressed air before the next sample was crushed. Samples that were crushed were either sawed half samples or whole samples (if unable to be sawed). The crushed samples were then weighed (typically 8.0 kg to 10.0 kg for a sawed sample) and subjected to a systematic splitting process using an industry standard riffle splitter. The original sample (typically 8 kg) was split into two 4 kg samples and they were, in turn, split into four 2 kg samples. Two of these (one from each of the 4 kg splits) were assigned to a reject bag and the remaining two were split to produce four 1 kg samples. Two of these (again, one from each of the original 4 kg splits) were then mixed to form the main assay sample. Where appropriate, the remaining two 1 kg samples were also mixed to make a primary crush duplicate.

The same splitting process was used for samples that were unable to be sawed, except a larger sample was retained since these samples were not stored in the core racks but with the primary crush duplicates. Those samples (main and duplicate) selected for analysis were then weighed, assigned their assay tags and sealed using a single-use cinch-tie. The samples were shipped to ACME in rice-sacks. Golder observed the 2005 and 2006 field sample preparation procedures and found them to be consistent with industry standards.

Illustrated on Figure 13-1 is a flow diagram of the field sample preparation procedures.


Figure 13-1 Adanac Field Sample Preparation Procedures

8.0 kg

4.0 kg 4.0 kg

2.0 kg

1.0

2.0 kg 2.0 kg 2.0 kg

1.0 1.0 1.0

2.0 kg 2.0 kgMain Sample, for Acme Labs.

Duplicate Sample, if requested.

6.0 kgor 4.0 kg

Reject for storage

Half Core8.0 kg

4.0 kg 4.0 kg

2.0 kg

1.0

2.0 kg 2.0 kg 2.0 kg

1.0 1.0 1.0

2.0 kg 2.0 kgMain Sample, for Acme Labs.

Duplicate Sample, if requested.

6.0 kgor 4.0 kg

Reject for storage

Half Core

*From Pinsent (2005)

13.1.2 Prior to Adanac

The sampling procedures used prior to Adanac’s 2004 to 2006 drilling programs are found in the Technical Report - Mineral Resource Estimate Ruby Creek Project (Blower, 2005).

• Prior to 1970, samples were prepared for assay in various laboratories. There was no preparation in the field.

• Pre-1979: Kerr Addison (1970), after constructing a bucking room and assay lab on site, completed both the field and laboratory sample preparation at the property.

• 1979 - 1980 Placer: Placer's field sample preparation procedures are virtually identical to Adanac’s 2004 protocols, except that Placer employed a second stage of crushing with a gyratory crusher to further reduce the 1/4" jaw crusher output to a -8 mesh sample weighing 1.75 lbs.


13.2 LABORATORY SAMPLE PREPARATION PROCEDURES

13.2.1 2004, 2005 and 2006 ADANAC

Samples from Adanac’s drilling programs from 2004 onwards were submitted to ACME Analytical Laboratories for molybdenum and trace element analysis. Check sampling and specific gravity estimates were submitted to ALS Chemex.

Samples collected during the drilling campaigns were prepared for molybdenum and trace element assaying. They were crushed to 70% passing 10 mesh and splits (250 g) were then pulverized to 95% passing -150 mesh.

Methods used for sample preparation prior to the 2004 and 2005 drilling program are provided in the Technical Report - Mineral Resource Estimate Ruby Creek Project (Blower, 2005).

13.2.2 1979 — 1980 Placer

Placer's 800 gram crushed samples were shipped to their laboratory, where the entire sample was pulverized, but the targeted pulp specifications are not known.

13.2.3 Pre-1979 Kerr Addison

Kerr Addison's laboratory sample preparation procedures in 1970 are shown in Figure 13-1. The sample preparation procedures prior to 1970 were not available.


Figure 13-2 Kerr Addison's Laboratory Sample Preparation Procedure

13.3 Analytical Procedures

13.3.1 Adanac 2004, 2005 and 2006

The molybdenum assaying method that was applied to all the 2004 to 2006 samples (including blanks, duplicates and standard samples) by ACME was the multi-element method. This method takes a 1.0 gram split pulverized sample first digested by aqua regia and then analyzes the resultant solution for molybdenum (Mo) using Inductively Coupled Plasma Emission Mass Spectrometry (ICP-MS).

Trace element assaying of the 2005 and 2006 samples by ACME was prepared the same as above and then a 0.25 gram sample was heated in HNO3-HCLO4-HF to fuming and taken to dryness. The residue sample was dissolved in HCL and the resultant solution was then analyzed for 41 elements (parts per million) using ICP-MS.

Note that the values returned are the total molybdenum content of the rock, as they combine the sulphide with any oxide molybdenum that may be present. However, in practice, previous work showed that there is little or no molybdenum present in oxide form. In Golder’s opinion, the


ICP method is the best practice for measuring the molybdenum content of molybdenite mineralization in rock.

13.3.2 1979 — 1980 Placer

Placer completed all of their analytical work at their in-house laboratory in Vancouver. The samples were analysed by Atomic Absorption Spectrometry (AAS) and were reported as % Mo in MoS2 (the sulphide portion of the total Mo content after removal of the oxide portion) (Christopher and Pinsent, 1982).

13.3.3 Pre-1979 Kerr Addison

Several laboratories were used for the pre-1979 drilling (Dagbert and David, 1975). Details on the 1970 analytical techniques are not available. Assaying in 1970 was completed by Kerr Addison on site in their own laboratory using a colorimetric - spectrophotometric procedure for most samples, and a gravimetric method for higher grade or mill product determinations (Chapman, Wood & Griswold, 1971). The on-site lab results were apparently checked with duplicate assays at Loring Laboratories.

No samples were available prior to the 2004 drilling program; therefore, no check assays have been completed on sample information prior to 2004. Adanac drilled five twin holes in 2004 to test the previous assaying results. The results from the twin hole drilling program are discussed in Golder’s 2006 Mineral Resource Estimate Report.

13.4 Quality Assurance and Quality Control

13.4.1 Adanac Procedures

Adanac has employed a comprehensive program of QA/QC consisting of inserted blanks (5%), standards (5-10%) and duplicate samples (5%) from the 2004 to the 2006 program and has employed the same program for the 2007 program. The QA/QC program employed follows the same QA/QC program as the one developed in 2004 which is described in the Mineral Resource Estimate Reports by Golder (2006 and 2007) and AMEC (Blower, 2005). There is no information available on QA/QC procedures or results from the drilling completed prior to 2004. Samples submitted to ACME included split drill core samples, blanks (two types) and standards (two types). Golder has reviewed the assaying results from the duplicates, blanks and standards that were provided by Adanac in 2005 and 2006. A description of 2006 procedures for the blanks, standards and duplicates is as follows and is provided in the July 23, 2007 Mineral Resource Update Report Update. Information regarding a review of the 2005 QA/QC program is


provided in the 2006 Technical Report – February 8, 2006 Mineral Resource Estimate Ruby Creek Molybdenum Project

2006 Blanks

Blank samples were inserted in the sample stream to monitor contamination. Adanac used two types of blanks: (1) bags of pre-crushed, commercially purchased, poultry grit quartzite (Blank), and (2) locally derived volcanic scoria from Ruby Mountain, west of Ruby Creek (Blank-S). The poultry grit blanks did not go through the on-site crusher and were inserted as a standard. However, those designated as Blank-S, composed of locally derived volcanic scoria from Ruby Mountain, did go through the crusher and they provide a check on sample to sample contamination during the crushing process. Approximately 5% of the samples submitted for a drill hole included blank samples with no molybdenum mineralization.

2006 Standard Reference Material

During the 2006 drilling program, two commercially purchased molybdenum standards were included as standard reference materials during the submission of drill core samples to ACME for both standard ICP-MS molybdenum assaying and trace element assaying. These standards included WCM Cu 111 and WCM Cu 132. The standard assaying results for WCM Cu 111 is 0.83% Cu, 0.117% Mo and 105 g/t Ag. The standard assaying results for WCM 132 is 0.17% Cu, 0.045% Mo and 27 g/t Ag and 0.17 g/t Au.

2006 Reject and Pulp Duplicates

During analysis, ACME also inserted reject duplicates from a second 250 g split from the original sample, and pulp duplicates from a second split from the original pulp, as requested by Adanac.

13.4.2 QA/QC Results –2006 Samples

The analytical methods employed on the 2006 samples from the Ruby Creek deposit by ACME were the same as those used during the 2004 and 2005 sampling programs. The ICP-MS analytical method used on the 2006 samples considered the best practice for measuring the molybdenum content in the drill core samples.

The assay results from the blanks, standards and duplicates from the 2006 QA/QC program were reviewed by Golder with the following results:

• 18 Blank samples were reviewed and had an assay range between 0.0005 and 0.004% Mo. Only one sample had an assay of 0.004% Mo with the remaining at or below 0.001% Mo.


• 27 Blank-S samples were reviewed and had an assay range between 0 (below detection) and 0.006% Mo.

• 39 WCM Cu 111 standard samples were reviewed and had a range of 0.103% Mo and 0.123% Mo with a mean and standard deviation of 0.111% Mo and 0.0049% Mo, respectively. Only 1 of the 39 samples exceeded 2 standard deviations.

• 39 WCM Cu 132 standard samples were reviewed and had a range of 0.0398 and 0.048% Mo with a mean and standard deviation of 0.0432% Mo and 0.002% Mo, respectively. Only 1 of the 39 samples exceeded 2 standard deviations.

• 172 duplicate samples were reviewed using Q-Q plots of assay 1 versus assay 2. In general, these plots showed reasonable agreement with some outliers occurring at all grades (both low and high).

Therefore, in general, there appears to be no contamination in the samples crushed on site since the Blank-S samples had a maximum assay of 0.006% Mo. Also, the standards and second blanks submitted show no obvious contamination and reasonable accuracy and precision with only 2 of the 78 standards reviewed exceeding 2 standard deviations above their recorded assay grade. The duplicate samples reviewed also showed reasonable agreement, which indicates reasonable precision and no contamination between samples. The review of the 2006 sampling program indicates that the samples were sufficiently accurate, free from contamination, precise, under control and were included in the 2006 Ruby Creek Datamine Database for the calculation of the July 23, 2007 Mineral Resource Update.


14.0 DATA VERIFICATION

The data verification checks that were completed on the drill hole data prior to them being used in the July 23, 2007 Mineral Resource Update is provided as follows.

1. A check of the drill hole data against original spreadsheet records in the database.

2. A review of the 2006 blanks, duplicate (Q-Q plots) and standards.

3. Site visit completed on August 22, 2006 to review core logging and sampling procedures. No independent samples were collected during the site visit, but visible molybdenum mineralization was observed and independent samples were collected in 2004 (Blower, 2005) and 2005 (Palmer, 2006).

Verification checks were completed on the 2005 and earlier data and are provided in the Technical Reports by Palmer (2006) and Blower (2005). The data verification checks completed on the 2006 data is discussed above with the exception of the drill hole co-ordinate translation from UTM NAD-27 to UTM NAD 83 and as described in the following sections.

Approximately 5% of the 2006 drill hole Excel spreadsheets were visually reviewed against the Datamine Database. The drill hole samples in the database prior to 2004 were provided as ASCII files. These ASCII files were provided from AMEC based on electronic spreadsheets provided by Adanac. These drill hole samples were added to the Ruby Creek Datamine Database created in 2006 and were visually checked against the 3D geological and mineralization models created.

No significant discrepancies were encountered during the check. The collar locations for drilling data from 2005 and earlier have undergone a conversion and translation from UTM NAD-27 to UTM NAD 83. Some drill hole samples (pre-2004 historical data) did not have information pertaining to percent recovery and main lithology identification, but were still included in the Mineral Resource Estimate since assay data was available. All drill holes with missing information (i.e. no Mo assay values for overburden samples) were flagged with a negative value (typically -2) and were not included in the mineral estimate.

A review of the 2006 sampling data including blanks, duplicates and standards was completed in previous sections.

Previous data verification checks completed in 2004 and 2005 and on earlier data are provided in the 2005 (AMEC) and 2006 and 2007 (Golder) Mineral Resource Estimation Reports.

Golder concludes that the assay and survey data used in the July 23, 2007 Mineral Resource Update were sufficiently free of error to be adequately used for resource estimation of the Ruby Creek Molybdenum deposit.


15.0 ADJACENT PROPERTIES

Adjacent properties are not relevant to the review of the Ruby Creek property.


16.0 MINERAL PROCESSING AND METALLURGICAL TESTING

A detailed description of the mineral processing and metallurgical testing is contained in the Process and Infrastructure Design and Cost Estimate report by Wardrop in 2006. A simplified flowsheet is shown in Figure 16-1.

Figure 16-1 Simplified Flowsheet

The process design is based on an average of 23,000 tpd mill operation. Average annual throughput would be 7.6 million tonnes per year.

16.1 Mineral Processing

The milling process is based on test work done by SGS-MinnovEX and adjusted by test work by G&T to account for a coarser primary grind. Three stages of crushing, ball milling, froth flotation, dewatering and drying are used to produce and package a high grade molybdenite concentrate (MoS2). Run.of.mine (ROM) ore is dumped into the gyratory crusher and crushed to


approximately -200mm. From there it is conveyed to the 130,000 tonne capacity coarse ore stock pile; of which approximately 40,000 tonnes is live capacity. From there it is reclaimed and conveyed to screens where the - 50 mm fraction – the finished product- goes to the HPGR units. Screen oversize is crushed in one of two cone crushers with cone crushed product recycled to the screens. (Secondary crushing is a closed circuit operation). The -50 mm product is fed to the high pressure grinding rolls in open circuit. A small amount of “edge recycle” of HPGR product is allowed for, but otherwise all product is fed directly into the ball mill. Ball mill product is classified through cyclones. The finished product has a P80 = 275 microns. Oversize material is recycled to the ball mill. The flotation comprises rougher/scavenger tank cells where rougher concentrate (2%) is reground in a tower mill to a much finer state (P80 = 20 microns). This product is cleaned in a series of stages; trace impurities are reduced to a minimum and the final molybdenum concentrate is a high grade product.

Ruby Creek mineralization contains minimal trace impurities which are pyrite, chalcopyrite, sphalerite, galena and bismuthinite. The extensive mineralogy work shows almost all are liberated in the grinding and regrinding stages and can be mostly removed in the cleaning stages with the uses of depressant D910 (sodium thiophosphate) in reasonable amounts. Other reagents for flotation are diesel (kerosene) and pine oil which are used at natural pH (7.8 – 8.2) in the milled pulp.

Final concentrates are thickened and settled with aid of flocculant and dewatered in a pressure belt press. Filter cake is dried in a rotary hearth drier, cooled in a storage bin and afterwards packaged in tote bags (~ 2 tonne capacity), weighed and moved to a storage area.

16.2 Metallurgical Testing

Kerr Addison contracted Britton Research to conduct bench scale and pilot plant test work (August 1969 to January 1971). The pilot plant work was done on site through a 100 ton/day plant which was fed from almost 10,000 tons of ore produced from extensive underground development which included muck from raises on specific drill holes. Almost 7000 tons were processed as several discrete lots from these raises and some lateral development. Results were consistent and well documented, and a basic flowsheet was established.

SGS-MinnovEX used their proprietary technologies Comminution Economic Evaluation Tool (CEET) and Flotation Economic Evaluation Tool (FLEET) to design the comminution and flotation process for Ruby Creek. Initially, 30 core samples were collected from the 2004 drilling and sent to SGS-MinnovEX in May of 2005. For the Feasibility Study, an additional 70 samples from the 2005 drilling were sent to SGS-MinnovEX. Adanac’s geologist, Dr. Robert Pinsent, ensured that these samples were widely spaced across the orebody and the lithology types and that they would represent fully the variability occurring within the deposit.


The test work was completed and the reports were issued to Adanac and Wardrop in February 2006.

In December 2005, at Adanac’s request, B.C. Mining Research Limited completed tests on alternative technologies for the regrinding options in the flotation cleaning stages and settling tests on molybdenite concentrate, rougher tailings, and reground concentrate.

In December 2006, at Adanac’s request, G&T completed pilot scale and locked cycle flotation studies and modal assessments of the cleaner circuit simulations.

Further details of both programs are provided in the following sections.

16.2.1 Kerr Addison’s Pilot Plant (Britton 1969 – 1971)

• 6251 tons of underground “ore” (average assay 0.12% Mo) were milled in a pilot plant. Range of feed grade was 0.056% Mo – 0.151% Mo, and sulphide molybdenum recoveries were in the range 94.6% - 97.1%.

• Good grade concentrates were produced (average was >57% Mo) and lead was the only impurity which exceeded the acceptable limit (0.06% or slightly higher)

16.2.2 SGS-MinnovEX Test Work

• Grinding tests showed the Ruby Creek samples to be uniform with only small variations in hardness. The ore is described as medium to soft with average work index of 12.5 kWh/tonne.

• Flotation tests confirmed expectations of being able to produce a final concentrate grade of 54% Mo grade and overall recovery of 89% - 90% from a mill feed of 0.084% Mo at a flotation feed size of P80 = 210 microns. The 0.084% Mo mill feed grade is in line with grade expectations for the first five years of operations.

• The flotation tests done using small samples often did not allow five stages of cleaning to be done because of insufficient production of concentrate as the cleaning stages advanced. SGS-MinnovEX simulated the flotation data by means of its proprietary FLEET algorithm which indicated that the flotation performance on an average mill feed (0.084% Mo) would yield a 54% Mo grade at a 90.1% recovery based on primary grind size of P80 = 180 microns. These data are conservative when compared to results from the Kerr Addison pilot plant work (1969 – 71).

• Mineralogical work done on SGS-MinnovEX’s behalf was superficial and brief and threw no light on Adanac’s concerns for the seemingly lower overall molybdenum recovery into a final concentrate.


• SGS-MinnovEX recommended a flash flotation stage in the grinding circuit to remove liberated MoS2 as soon as possible but were not able to simulate the effect this would likely have on overall molybdenum recovery.

• Very little gangue floats by attachment to the molybdenite into the rougher concentrate. However, SGS-MinnovEX concluded satisfactory liberation still requires two stages of regrinding, each to achieve P80’s of about 40 and 70 microns within staged cleaning. (Regrind sizes are similar because cleaning removes the gangue leaving molybdenite concentrated into the coarse fraction.

16.2.3 B.C. Mining Research Limited Test Work

Fine grinding tests were conducted at the University of British Columbia to assess alternative technologies for regrinding the rougher flotation concentrate using either a Netzch (ISA) mill or a Stirred Media Detritor (SMD). The testing shows:

• Either mill, ISA or SMD, produced similar results. A comparison of different grinding media indicated that the 2 mm ceramic media resulted in lowest power requirements.

• Using 2 mm ceramic media, the power input required to grind from a P80 of 200 microns to a P80 of 40 microns is about 30 kWh/t.

• Molybdenite concentrate required high-energy (about 200 kWh/t) to grind from a P80 of 31 microns to a P80 of below 20 microns using either mill.

BC Mining Research noted stirred mills are usually operated in open circuit and they produce a product with a narrow particle size distribution. At a P80 of 40 microns, the size distribution would be similar to that produced with a ball mill in closed circuit with a classifying cyclone, but at finer grind sizes stirred mills may produce a narrower size distribution which benefits flotation and dewatering. The high stress intensities in stirred mills effect particle breakage.

16.2.4 G&T Test Work

A programme of tests to confirm expectations from SGS-MinnovEX’s work backed by detailed mineralogy (including modal analyses) was done in the latter part of 2006. The work carried out at G & T Metallurgical Services revealed that acceptable Mo concentrate grades and recoveries are achievable at a coarser primary grind sizings of about 275 µm K80. This was in contrast to simulated results from the SGS-MinnovEx work carried out on smaller samples. The projected optimal primary grind sizing from the SGS-MinnovEX work suggested a finer optimal primary grind sizing. The mineralogical data proved to be the key. Modal analyses presented details confirming the degrees of mineral liberation and how these affected flotation performance. In all cases (SGS-MinnovEX tests and G & T’s tests) every effort has been made to ensure full representation of material being tested.


• G & T confirmed the need for only one stage of regrinding (to a P80 + 20 microns)

• High molybdenite recovery into rougher concentrates attained with low mass pull (<3%).

• Only three to five stages of cleaning required to produce high grade and high recovery of molybdenite into final concentrate.

• Only lead noted to be an impurity which may have to be leached to achieve acceptable limit.

The comprehensive testing done during 2005 – 2006 produced results which corroborate Kerr Addison’s pilot plant results (1969 – 1971). High recovery of molybdenite into a high grade concentrate containing minimal impurities should be expected from a full scale production operation. Overall molybdenum recoveries in excess of 90% achieved in test work have not been considered in the project economic modelling.


17.0 MINERAL RESOURCE AND MINERAL RESERVE ESTIMATE

The July 23, 2007 Mineral Resource Update for the Ruby Creek Molybdenum Project was calculated under the direction of Paul Palmer, P.Geo., P.Eng., of Golder Associates Ltd. as the Qualified Person. Richard Gaze and Dr. Sia Khosrowshahi of Golder Associates Pty. assisted with the geostatistical analysis and Mineral Resource Estimate under the direction of Paul Palmer. This is the second time Paul Palmer has worked for and reported a mineral resource estimate for the Project.

The July 23, 2007 Mineral Resource Update was based on a 3D geological and block model constructed with commercial mine planning software (Datamine and Vulcan) based on the 2006 Datamine Database which includes all data collected from 2006 and earlier. No 2007 drilling information is included in the July 23, 2007 Mineral Resource Update.

The Project limits were 588396 to 590896 East, 6619174 to 6621674 North and 700 m to 1900 m elevation (NAD-83 UTM co-ordinates). The previous mineral resource estimates used drill hole collars in the NAD-27 UTM co-ordinate system. The July 23, 2007 Mineral Resource Update used drill holes collars in the NAD-83 UTM co-ordinate system. This initially required transferring all the drill hole data (i.e. collar information) and all the previously created 3D geological and mineralization geometries, and generating a new block model based on the NAD-83 UTM co-ordinate system. The block model used in the July 23, 2007 Mineral Resource Update was not rotated, measured 20 m east by 20 m north by 12 m (RL-elevation high) and sub-celled to a minimum of 5 m by 5m by 3 m.

The Ordinary Kriging (OK) interpolation method was used for resource estimation of %Mo using variogram parameters defined from the geostatistical analysis that was completed for the 2006 Mineral Resource Estimate. In addition, new variogram parameters were identified from the structural data that was collected from the 2006 inclined drilling program and optical televiewer analysis outlined in previous sections. Based on this analysis, the new variogram search parameters were changed to a vertical direction search as opposed to a horizontal direction search. The new variogram search parameters were only applied to the central deposit area, which is where the highest drill density and underground data is located. The delineation was accomplished by creating a new 3D mineralized domain (Zone 60) in the central pit area such that the vertical variogram search parameters were applied to Zone 60 and horizontal search parameters were applied to the mineralized domains outside of Zone 60 (partial Zone 6 and all of Zone 7).

In addition, the 2006 drilling data (16 bore holes) were applied to the current 3D mineralization geometries and modified to reflect the new data including slight interpretation modifications in other area. Based on the updated 3D mineralization geometries, two mineral resource estimates


were completed. The first resource estimate (Model 1) is based on using the geostatistical analysis defined from the 2005 and 2006 Mineral Resource Estimates, which used a variography orientation that was horizontal. The second resource estimate (Model 2) is based on using both vertical variography orientation for Zone 60 and horizontal variography orientation for Zones 6 and 7, outside of Zone 60.

Mineral resource estimates for measured, indicated and inferred resources were calculated for each model using cut-off grades between 0.02 and 0.10 %Mo. The comparison of the tonnage and grade estimates from each model indicated that the combined measured and indicated tonnage for Model 1 was slightly higher (approximately 0.5% higher) when compared to Model 2 and the grades were the same. Therefore, Model 1 was considered still the most appropriate method in estimating the mineral resource for the Project and was used in the July 23, 2007 Mineral Resource Update. The Model 2 mineral resource estimate supports the current estimates and should be reviewed again if additional inclined drilling data is collected.

17.1 The Database

The database that was used in the July 23, 2007 Mineral Resource Update was constructed initially in Datamine and was completed on drill hole information with the exception of historical sample data from underground development. The majority of the data (approximately 60%) is historical drill hole information (Kerr Addison, Climax and Placer) that has now been updated with drilling programs by Adanac in 2004, 2005 and 2006 and includes a total of 71 drill holes. The Datamine Database used in the mineral resource estimate is composed of 266 drill holes and 15,328 samples and is summarized in Table 17.1.

Table 17.1 Drill Hole Data in the July 23, 2007 Mineral Resource Update

Campaign Years Drill Hole Prefix

No. of Drill Holes

No. of Samples

Percentage of Database (%)

Kerr Addison (Adanac Mining and Exploration, John’s Manville and Kerr Addison)

1966 – 1972 KA 105+17

pseudo -drill holes

6,073 39.5

Climax Moly 1973 CL (100 Series 7 365 2.4 Placer Development 1979 – 1980 PD (200 Series) 66 3,145 20.5 Adanac Gold Corporation (Adanac Moly Corp.) 2004 AD (300 Series) 36 2,886 18.8

Adanac Moly Corporation 2005 AD (300 Series) 19 1,621 10.6 Adanac Moly Corporation 2006 AD (300 Series) 16 1,238 8.1

Total 266 15,328 100.0


The database is comprised of vertical, inclined drill holes and underground sampling data (pseudo-horizontal drill holes). Each sample in the database has the following attribute information: hole ID, from and to sample distance, total sample length, %Mo, lithological identification (primary and secondary), sample number, percentage recovery and RQD. Samples in the overburden were generally not populated with %Mo, recovery and RQD and some earlier samples did not collect RQD information. Sample lengths in the database were on average 10 ft long (3.05 m). Any data not available was either left as blank in the original database or later flagged with a negative code (-9 or -99).

Historical drill hole data in the Datamine Database have been outlined in previous sections of this report, with additional details provided in the Technical reports by Blower (2005) and Palmer (2006).

17.2 The Geological Model

The geological model that was generated for the Ruby Creek deposit comprised 3D wireframe geometries of geological interpretations and a 3D block model created in Datamine and Vulcan software. Six 3D geometries were created: three represented the main lithological units (primary) and three represented the main molybdenum mineralization. One additional mineralization geometry (Zone 60) was included in the July 23, 2007 Mineral Resource Update, which was located in the central pit area. Two surfaces (DTM) were also used to represent the overburden/bedrock interface and the topographic surface. The 3D block model was composed of a 20 m (x-easting) by 20 m (y-northing) by 12 m (z-elevation) parent blocks that were further subdivided into a minimum of 5 m (x) by 5 m (y) by 3 m (z) sub-celled blocks.

The 3D geometries used in the July 23, 2007 Mineral Resource Update were updated from the January 2006 3D geometries based on the new information from the 2006 drilling program (16 bore holes) and slight interpretation modifications in other areas.

The 2007 3D geometries were used for coding both drill hole data and the block model that defined the spatial zones for the estimation of grades and bulk density assignment in the 2007 resource model (Model 1). The naming convention used for the 2007 2D and 3D geometries is illustrated on Figure 17-1 (Local Grid 00) and summarized in Table 17.2.


Figure 17-1 Section View of 3D Mineralization Zones and Rock Types for Ruby Creek Deposit


Table 17.2 2007 Ruby Creek Molybdenum Project 2D and 3D Geometries

Field Value 2D and 3D Wireframe Geometry Description

Rzone 1 ob_nov7.00t Overburden Surface Rzone 2 RZN2.00t Rock Type 2 Rzone 3 RZN3.00t Rock Type 3 Rzone 4 RZN4.00t Rock Type 4 Zone 6 MIN_ZONE.00t Mineralization Zone 6 Zone 7 zone7_nov7.00t Mineralization Zone 7 Zone 60 Mineralization Zone 60

Rzone 19 topo_clipSept27.00t Topographic Surface Zone -99 Background Mineralization

Rzone 1 represents the geometry between the bedrock and the overburden. Rzones 2, 3 and 4 represent a simplified 3D geometry interpretation of the rock types in the Ruby Creek deposit. The lithological units that were classified inside Rzones 2, 3 and 4 are summarized in Table 17.3. The rock types in Rzones 2, 3 and 4 are a simplification of the various lithological units in the Ruby Creek deposit which, for resource modelling purposes, were acceptable to combine in order to define densities and tonnages in the block model. Rzones 2 and 3 rock types are geographically located south of the Adera Fault and Rzone 4 rock types are located geographically north of the Adera Fault.

Table 17.3 Primary and Secondary Lithological Units

Text Code Rock Type Code (RZone) Description

CGQM 2,4 Coarse-Grained Quartz Monzonite CGQM-T 2,4 CGQM – Transition Variety

CQFP 2,4 Crowded Quartz Feldspar Porphyry SQFP 2,4 Sparse Quartz Feldspar Porphyry

CGQM-H 2,4 CGQM – hybrid MQMP 2,4 Mafic Quartz Monzonite Porphyry SQMP 3 Sparse Quartz Monzonite Porphyry CQMP 3 Crowded Quartz Monzonite Porphyry FGQM 2,4 Fine-Grained Quartz Monzonite MFP 2,4 Mafic Feldspar Porphyry BSLT 2,3 Basalt


Zones 6, 7 and 60 represent the 3D geometry interpretations of the main molybdenum mineralization in the Ruby Creek deposit. The criteria used to classify drill hole samples inside the mineralization zones are summarized as follows:

• %Mo greater than or equal to 0.04%;

• primary or secondary rock identification as either FGQM or aplite dykes; or

• presence of silicification alteration.

The 3D poly lines corresponding to mineralization zones were modified based on the 2006 drilling and only impacted Zones 6 and 60. Considerations were also made for continuity between sections especially where data was not available. The 3D poly lines were then constructed into 3D geometries and validated in Datamine and Vulcan software. Representative sections are illustrated in Appendix A.

Zones 6 and 60 represent the main mineralization located south of the Adera Fault. Zone 60 (inside of Zone 6) also represents the main mineralization in the central area of the deposit in which there is a greater density of sub-horizontal and sub-vertical veining and potentially a breccia/feeder to the deposit. The breccia/feeder zone is located toward the centre of the deposit and was reported to intersect the underground development sampling completed by Kerr Addison. The smaller Zone 7 mineralization geometry is dominated by sub-horizontal veining and molybdenum mineralization located north of the Adera Fault.

Rzone 19 represented the 2D topographic surface over the Ruby Creek deposit. All Mo mineralization outside of Zones 6, 60 and 7 were coded as Zone -99 and represented the low grade (%Mo less than 0.04%) background mineralization. Some mineralization pockets with %Mo greater than 0.04 did occur outside of Zones 6, 60 and 7, but were not included inside these geometries because they were localized high grade pods with no sample data to confirm continuity from section to section.

17.3 Wireframe Validation

Prior to block model generation and populating of the block model, the Datamine 3D geometries were validated in the Vulcan Software using standard wireframe validation check routines and by slicing sections through the individual wireframes for comparisons against the drill hole database. The process verified the following:

• consistency, by testing triangle edges;

• correct solid triangulation closure;

• self-intersection, by testing for self-crossing triangles; and


• correct spatial location of individual points within the wireframe.

17.4 Data Preparation and Compositing

Drill hole data for the Ruby Creek deposit was predominantly sampled on 10 ft lengths (3.05 m). The global distribution of raw sample lengths is shown as a cumulative log-probability plot, which is illustrated on Figure 17-2 from Palmer (2006). This plot illustrates that 95% of the data is represented by the 3.05 m length.

Figure 17-2 2006 Global Distribution of Raw Sample Lengths (Palmer, 2006)

Considering the nominal most frequent raw sample interval (approximately 3 m) and the likely vertical mining selectivity for the project, a 6 m down hole compositing interval was selected. This length was a multiple of the nominal approximate 3 m interval and a divisor of the anticipated mining bench height of 12 m, and the block model vertical dimension of 12 m.

Prior to compositing, the raw sample intervals were generated from the Datamine database and flagged to the 3D geometries listed in Table 17.3 and assigned codes for mineralization zone and rock type. The flagged intervals were then uploaded to a new flagged database.


From this flagged database, a 6 m composite file was generated for statistical and geostatistical analysis using run-length down hole compositing. All composites were broken at the 3D mineralized zone contacts (Zones 6, 60 and 7) resulting in some composite lengths of less than 6 m.

Any gaps in the down hole sequence representing unsampled intervals were excluded from the compositing process, with no default values assigned to represent the missing sample intervals. Due to the use of a length-weighting approach in the interpolation method used for resource estimation, residual composites arising from the breaking of 6 m compositing at geological contacts were not removed from the composite dataset.

No filtering of raw data on the basis of the sample recovery was carried out, mainly due to large amounts of missing values in the recovery data. No relationship between sample recovery and %Mo was evident from a statistical analysis.

17.5 Declustering

The exploration drill holes in the Ruby Creek deposit were on a nominal pattern of 90 m by 90 m (approximately 300 ft by 300 ft) spaced drill holes, with in-fill drilling in the central region of the resource down to 30 m by 30 m (approximately 90 ft by 90 ft) spacing. A location plan of the drilling is shown on Figure 11-1.

Due to the clustered nature of the drill holes, spatial declustering was carried out using a 90 m by 90 m by 12 m moving window in order to achieve more representative global statistics. A standard cell declustering algorithm was applied to determine the declustering weights.

17.6 Spatial Trend Analysis

Spatial trend analysis was carried out independent of domain contacts in 2006 to assess grade trends in %Mo content across the resource and is provided in Palmer (2006). This was done using fixed block averages, where 6 m composite data was averaged into 90 m by 90 m by 12 m blocks across the resource, and the mean block grades displayed spatially. This provides an indication of whether there are particular grade orientations or zonation of elevated grade that need to be accounted for in the resource estimation. The spatial trend analysis showed some indication of a South-Easterly grade trend.

17.7 High-Grade Treatment

High-grade treatment for estimation was applied using a spatial restraining method to avoid over-estimating the grade of %Mo in the resource domains. In this method, samples above a


designated threshold are flagged to individual blocks in the model and their spatial influence restricted to a single block. Therefore, instead of capping high grade values, they were restricted spatially.

Preliminary high-grade thresholds were selected from an examination of cumulative log-probability plots for each domain, and refined through estimation validations of the reproduction of the estimated mean domain grade in the Kriging. After the preliminary estimate and model validations were run, high-grade thresholds were adjusted where required, and the model re-estimated.

A summary of the high grade thresholds implemented in the grade estimation is provided in Table 17.4.

Table 17.4 High-Grade Thresholds for %Mo by Zone

Mineralized Zone %Mo High-Grade Threshold

Approx. Distribution Percentile

6 and 60 0.3 99.5 7 0.2 97.5

-99 0.15 99.7

17.8 Variogram Analysis

17.8.1 Variography Objectives and Approach

The objectives of the variographic analysis were the following:

• to establish the directions of major grade continuity for each element in the domains; and • to provide variogram model parameters for use in geostatistical grade interpolation.

The variographic analysis was competed using in-house Golder software and was based on the 6 m composited %Mo data for the combined Zones 6, 60 and 7. Conventional 3D directional variography was used for spatial continuity analysis using 3D conical search methodology. Spherical scheme models were used for modelling of the major, semi-major and minor orthogonal directions of continuity. These models were generated during the 2006 Mineral Resource Estimate and are provided in Palmer (2006). The variogram parameters used in the February 22, July 23, 2007 Mineral Resource Update were taken directly from the 2006 analysis in Palmer (2006). In addition, a new variogram analysis was used, based on the structural data from the 2006 inclined drilling program. The new variography analysis was applied to the mineral Zone 60 to determine if using the new variogram parameters better defined the grade of the mineral


resource estimate in this area. The results from the new variogram analysis were not significantly different than the 2005 variogram analysis. The grades from Zone 60 were not significantly different; therefore, the 2005 variogram analysis (horizontal variogram) from 2006 was used in the February 22, 2007 Mineral Resource Estimate.

17.8.2 Summary of Variography Parameters

Table 17-5 provides a summary of variography model parameters and orientations modelled for the combined mineralized Zones 6, 7 and 60, determined in the 2006 and 2007 Mineral Resource Estimates.

Linear orientations are defined in terms of plunge and plunge direction (i.e. plunge plunge direction). Resultant planes are defined in terms of dip and dip direction (i.e. dip dip direction).

The ‘separation’ is the angle between the major and semi-major vectors. The ‘rotation’ applies to the semi-major axis; it is the angle required to rotate the search ellipsoid around the major axis vector to align the ellipsoid with the interpreted resultant plane for grade interpolation.

Table 17.5 Variography parameters for Zones 6, 60 and 7

Correlogram Parameters for %Mo Grades Structure 1 Structure 2

Axis Direction Nugget Diff. Sill Range Diff. Sill Range Nugget/Sill Major Axis 0.300 0.340 42.50 0.320 125.00 31

Semi-Major Axis 0.300 0.200 39.00 0.580 137.00 28 Minor Axis 0.300 0.510 15.50 0.170 98.50 31

Major Axis Semi-Major Axis Resultant Plane Major & Semi-

Major Rotation Angles Mineralization

Zone Plunge/Direction Plunge/Direction Dip/Dip

Direction Separation Angle R1/R2/R3

6, 60 and 7 0 140 0 050 0 050 90 140/0/0

KEY: R1/R2/R3 Search orientations (conventional left hand rule) R1 = azimuth rotation clockwise from north R2 = plunge along R1 direction (+ve = up, -ve = down) R3 = dip rotation around R1-R2 axis (+ve = anticlockwise, -ve = clockwise}

17.9 Update Block Model Parameters

A block model was constructed to cover the mineralization in the Ruby Creek deposit and sufficient surrounding waste for inclusion into open pit designs. The dimensions of the model are


provided in Table 17.6 and illustrated including mineralization Zone 6 (green) and 7 (yellow) on Figure 17-3.

Table 17.6 Block Model Dimensions for Ruby Creek Resource Model

Minimum Maximum Block size (m)

No. of blocks

Sub-block size (m)

Easting (X) 588 396 590 896 20 125 5.0

Northing (Y) 6 619 174 6 621674 20 125 5.0

RL (Z) 700 1900 12 100 3.0

Figure 17-3 Plan View of Block Model Geometry

Blocks were filled inside each of the wireframed geological interpretations in a priority sequence to create the geological zones.

Two separate variables were established for mineralized zone (Zones 6, 60, 7 and -99) and rock type (Rzones 1, 2, 3, 4 and 19). Grade interpolation was constrained to the mineralized zone interpretations. Rock type was used for applying the density and corresponding tonnage.


Sub-blocks were used in each model to define the geological zone boundaries and the topography. Sub-blocks were estimated to the parent cell to help achieve acceptable local estimation quality.

17.10 Grade Interpolation

17.10.1 Grade Interpolation Methods and Objectives

The OK interpolation method was used for the resource estimation of %Mo for the Ruby Creek Molybdenum Project using variography parameters defined from the geostatistical analysis.

The resource estimation reflects an in situ resource estimate based on a nominal block selectivity of 20 m by 20 m by 12 m. No allowance has been incorporated for higher SMU selectivity or ore loss and dilution assumptions.

17.10.2 Ordinary Kriging Plan

The Kriging plan parameters used for grade interpolation %Mo are summarised in Table 17.7 with details provided as follows:

• All mineralized zones were estimated individually with their own data (hard boundary conditions). Waste zones were estimated with parameters from the mineralized zone.

• Sample weights were determined by the variogram model parameters (kriging approach).

• Block discretisation was set to 5 (x) by 5 (y) by 3 (z) to estimate block grades of 20 m by 20 m by 12 m parent blocks. Estimation of sub-cells in the model was performed to the parent cell size, so sub-cells received the parent cell estimate of 20 m by 20 m by 12 m.

• A maximum of four samples per discretised block was used, which equates to a maximum of 32 samples per estimate.

• OK was applied in three passes. Pass 1 used a search of 70 m by 70 m by 12 m. Pass 2 used a search of 140 m by 140 m by 24 m to estimate blocks not estimated in Pass 1, or blocks estimated in Pass 1 with less than two drill holes. Pass 3 used a 280 m by 280 m by 24 m to estimate blocks not estimated in the first two passes, but using a minimum of one sample. Pass 3 was designed to fill any remaining gaps in the model. Pass 3 in the waste zone (zone=-99) was set to the same dimension as the second pass, but with a minimum of two samples to limit grade extrapolation.

• Estimation was weighted by the sample length to account for variations in sample length due to the compositing process.


• High grades above the thresholds were spatially restrained to a single block for grade estimation to control the influence of the high grades. Therefore, high-grade values were restricted to the 20 m by 20 m by 12 m parent block.

Table 17.7 Ruby Creek Deposit Kriging Plan Parameters

Estimation Method Ordinary Kriging Search radius (pass 1/2/3)

X 70/70/12 Y 140/140/24 Z 280/280/24

Anisotropy (OK) X Defined by Variogram Y Defined by Variogram Z Defined by Variogram

Discretisation X 5 Y 5 Z 2

Search type Octant Minimum No. samples 4/4/1

Maximum No. samples per octant 4

17.11 Density Assignment

Adanac supplied Golder with the results of specific gravity samples that were analyzed by ALS Chemex. In all, there is approximately 1,000 samples collected from the 2004 to 2006 drilling programs. A statistical review was completed on the specific gravity samples (Palmer, 2006) based on lithological unit and rock type zone (Rzones 2, 3 or 4). The minimum, maximum and mean specific gravity estimates are summarized in Table 17.8. The mean bulk densities were assigned to all blocks in the model based on rock type using the values in Table 17.8.


Table 17.8 Rock Type Bulk Density Assigned to the Block Model

Rock type Min Bulk Density

Max Bulk Density

Mean Bulk Density

Default rock type 2.57

Rzone 2 2.19 2.88 2.57

Rzone 3 2.22 2.63 2.57

Rzone 4 2.37 2.71 2.55

All the specific gravity samples for the current Datamine Database (approximately 22% of database) were based completely on the 2004 - 2006 drilling samples. No specific gravity information was available from the historical data. Therefore, a representative amount of specific gravity samples have been collected from the 2004 - 2006 drilling programs. Additional samples should be collected from any future drilling programs to continue to expand the dataset, specifically in areas where no specific gravity sample data has been collected.

17.12 Mineral Resource Classification

Classification of the resource estimate was based principally on sample data density and geological confidence criteria. Consideration was also given to the classification based on the two previous mineral resource estimates (Blower, 2005 and Palmer, 2006) in order to achieve some continuity in terms of changes in proportions of resource categories with the additional drilling.

The previous mineral resource estimates used a confidence interval scheme for classification as follows (Blower, 2005):

• Measured Resource: estimates within +/- 15% at 90% confidence for quarterly production

• Indicated Resource: estimates within +/- 15% at 90% confidence for annual production

The application of this method indicated that the following drill densities were applicable with respect to resource categories:

• Measured Resource: drill hole spacing of 30 by 30 m or less

• Indicated Resource: drill hole spacing of >30 by 30 <90 by 90 m

• Inferred Resource: drill hole spacing of greater than 90 by 90 m.


Based on the mineralization style at Ruby Creek and the continuity identified from the variography studies, the above drill hole spacings were considered an acceptable method for resource classification of the Ruby Creek deposit and were used in the July 23, 2007 Mineral Resource Update.

The July 23, 2007 Mineral Resource Update included resource classification by digitising polygons on cross-section for both Indicated Resources (drilling <= 90 by 90 m) and Measured Resources (drilling <= 30 by 30 m) to better define the initial classification boundaries.

The following conditions were also applied to the mineral resource estimate (using Vulcan scripts):

• Initially, all blocks were set to an Inferred Resource;

• Blocks inside the indicated 3D geometry were set to Indicated Resource;

• Blocks inside the measured 3D geometry were set to Measured Resource;

• Measured blocks were downgraded to an Indicated Resource if the average weighted interpolation distance was greater than or equal to 50 m or the number of drill holes used to estimate the block was less than four;

• Measured blocks initially defined by the measured 3D geometry but located outside the main mineralization zones (Zones 6, 60 and 7) were downgraded to an Indicated Resource; and

• All blocks estimated with less than two drill holes per block were downgraded to Inferred Resource.

The following resource classification codes were used in the block model:

• Measured Mineral Resource (class = 1);

• Indicated Mineral Resource (class = 2); and

• Inferred Mineral Resource (class = 3).

17.13 Mineral Resource Summary

The Ruby Creek Mineral Resource Estimate is summarised in Table 17.9 reported at %Mo cut-offs from 0.02 to 0.10% (resource tabulation convention is greater than or equal to the cut-off).


Table 17.9 July 23, 2007 Mineral Resource Update

Resource Category Cut-off (%Mo) Tonnage %Mo Mo lb

0.020 55,831,000 0.068 83,698,000

0.030 54,300,000 0.069 82,600,000

0.040 49,106,000 0.073 79,029,000

0.050 41,389,000 0.078 71,172,000

0.060 30,151,000 0.086 57,165,000

0.070 21,909,000 0.094 45,403,000

0.080 14,556,000 0.104 33,374,000

0.090 10,411,000 0.112 25,706,000

Measured

0.100 6,612,500 0.122 17,785,000

0.020 387,278,000 0.042 358,593,000

0.030 238,954,000 0.052 273,935,000

0.040 163,801,000 0.060 216,669,000

0.050 109,444,000 0.067 161,658,000

0.060 61,471,000 0.077 104,350,000

0.070 37,664,000 0.084 69,749,000

0.080 18,813,000 0.094 38,987,000

0.090 9,848,100 0.102 22,145,000

Indicated

0.100 4,286,900 0.113 10,680,000

0.020 443,108,000 0.045 442,290,000

0.030 293,254,000 0.055 356,535,000

0.040 212,907,000 0.063 295,699,000

0.050 150,834,000 0.070 232,831,000

0.060 91,621,000 0.080 161,513,000

0.070 59,573,000 0.088 115,151,000

0.080 33,369,000 0.098 72,360,000

0.090 20,259,000 0.107 47,851,000

Measured +

Indicated

0.100 10,899,000 0.118 28,464,000

0.020 135,737,000 0.032 95,759,000

0.030 48,456,000 0.045 48,072,000

0.040 24,973,000 0.054 29,730,000

0.050 11,631,000 0.064 16,411,000

0.060 5,194,300 0.077 8,817,500

0.070 2,626,200 0.089 5,152,900

0.080 1,239,700 0.103 2,815,000

0.090 821,310 0.113 2,046,000

Inferred

0.100 493,790 0.125 1,360,800 Cut-off %Mo grades were classified as greater than or equal to and range from 0.01 to 0.10 in increments of 0.01. The mineral resource estimate has been completed in accordance with CIM standards of Estimation of Mineral Resources and Mineral Reserves.


17.14 Mineral Reserve Estimate and Mine Design

The final depth of the proposed open pit is planned to be at an elevation of 1,228 m. The North West rim of the pit will reach Molly Lake (El. 1,600 m). Molly Lake will be drained before the open pit mining is begun.

Open pit mining will initially be performed using a diesel-powered rotary drill and a hydraulic shovel with a 22.0 m3 as the primary loading tool with two 12.0 m3 front end loaders (FELs) as secondary loading tools. A fleet of six 150 ton capacity end-dump haul trucks will be used for ore transport. Road, dump and pit bench maintenance will be carried out with a fleet of compatible support equipment. After grid electric power becomes available, a second rotary drill, (electric) will be added to the fleet, as well as 22.0 m3 electric front shovel and seven additional 150 ton capacity end-dump haul trucks. The rock will be drilled, blasted, loaded and hauled to either the primary crusher, in the case of ore or designated locations for stockpile ore and waste rock.

The comminution circuit includes a single gyratory crusher followed by two secondary cone crushers in closed circuit with a screening plant. Cone crusher discharge feeds two high pressure grinding rolls (HPGRs) and a single ball mill designed to produce material for flotation. The flotation circuit includes rougher/scavenger and cleaner stages. To increase the degree of liberation and improve overall molybdenum recovery, a single regrinding stage is included within the cleaning section. The final concentrate is thickened, filtered, dried, packaged and shipped offsite for roasting. Tailings will be impounded and water recycled into the process.

A detailed plan for the property is currently being developed. This plan includes a construction camp, fresh water supply and sewage collections/treatment, power generation and distribution, fuel storage, communications, tailings impoundment, waste dump sites and accommodation for operating personnel, has been developed as an integral part of project development.

Construction is currently underway and the planned start up is scheduled in the year 2009. A construction work force is estimated to peak at 700 people. A summary schedule is provided in Figure 17-4.


Figure 17-4 Summary Schedule

The Golder Mine Feasibility Report completed in March 2006 served the basis of parameters for the detailed mine design work described in this report. For example, the geotechnical work and analysis required for the open pit design were undertaken by Golder in the Summer of 2005 and results are contained in their feasibility report.

Open pit optimizations were completed using a range of molybdenum prices from US$5.00 per pound to US$10.00 per pound. Previous work used molybdenum prices up to US$13.00 per pound to guide the location of mine facilities and related surface infrastructure. The open pit optimizations were performed with overall angles of 45 - 48.5° based on the geotechnical evaluation.

A molybdenum base price of US$10.00 per pound was used for all tonnage calculations and cut-off grade determinations. At this price, the calculated mining cut-off grade was 0.040% Mo and the processing cut-off grade (internal cut-off) was 0.030% Mo. The mining cut-off was set at 0.040% Mo and low grade material grading between 0.030% and 0.040% stockpiled for processing later.


A series of four phases or push-backs have been designed for the ultimate pit . The revised open pit design extends south southwest to incorporate reserves of the phase 5 push back identified in the Golder March 2006 Feasibility Report.

Phase 1 focused on the near surface higher-grade portion to maximize grade and subsequent early cash flows, thereby minimizing the payback period, increasing internal rate of return reducing other financial risks.

All pit phase designs were designed to accommodate a truck size of 200 tonnes capacity (which allows for a possible upgrade from the currently contemplated fleet of 150 tonne trucks, later in the mine life), utilizing 30 metre ramp width. The mining schedule uses a cut-off grade of 0.060% Mo for Phase 1 and 2 to generate a payback pit and speed repayment of mine capital. The remaining phase 3 and 4 were completed with a mining cut-off of 0.040% Mo and an internal cut-off grade of 0.030% Mo for the low-grade stockpile throughout. The ultimate pit design can be seen in Figure 17-5.


Figure 17-5 Ultimate Pit Design

The production schedule provides for 20 years of mining and 21 years of processing at a rate of 7.6 million tonnes per year. Based on the mining cut-off, production plan and schedule a Mineral Reserve was determined. This mineral reserve is presented in Table 17-10, and is estimated at 157.6 million tonnes grading 0.058% Mo.


Mining dilution for the Ruby Creek deposit was assessed at 5.1% along with an accompanying ore loss of 3% for a net tonnage gain of 1.9% above the undiluted tonnage. Undiluted grades were subsequently reduced with a 0.020% Mo diluting grade for the 5.1% dilution tonnage. Table 17.10 illustrates the reserve distribution by phase and classification.

Table 17.10 November 22, 2007 Mineral Reserve Estimate

Ore to Mill Stockpile Ore Proven Probable Proven Probable Total Tonnes %Mo Tonnes %Mo Tonnes %Mo Tonnes %Mo Tonnes %Mo

Phase 1 19,455,000 0.089 3,065,000 0.082 6,081,000 0.049 2,608,000 0.042 31,209,000 0.077 Phase 2 3,903,000 0.070 6,819,000 0.075 4,996,000 0.049 7,519,000 0.046 23,237,000 0.059 Phase 3 20,250,000 0.056 48,463,000 0.050 185,000 0.027 2,893,000 0.026 71,791,000 0.051 Phase 4 271,000 0.049 29,166,000 0.056 12,000 0.027 1,999,000 0.026 31,448,000 0.054 Total 43,879,000 0.072 87,513,000 0.055 11,274,000 0.049 15,019,000 0.039 157,685,000 0.058

A production schedule was developed, that ensures mill feed continuity during and between mining phases. Table 17.11 shows the mining and processing schedule by year.


Table 17.11 Ruby Creek Production Schedule


Table 17.11 Ruby Creek Production Schedule (Con’t)


Figure 17-6 and 17-7 present the actual phase (pushback) development sequence in plan and section, respectively. In general terms, mining proceeds from the highest value ore to the lowest value ore. The phase 1 pit is located in the northeast part of the deposit. Phase 2 ,3 and 4 expand the open pit to the south and southwest and deepen the northeast lobe to its ultimate depth.

Specific protocols have been included to account for the identification, handling and disposal of potentially acid generating waste.

Figure 17-6 Phase Pushback Sequence in Plan


Figure 17-7 Phase Pushback Sequence in Plan

The expected mine manpower has been detailed in Table 1712. Four departments fall under the Mine Operations area, as follows:

• Maintenance;

• Operations; and

• Engineering/Geology.

The total number of positions for the initial five years is estimated to be 115 people. This allows for a more flexible rotation schedule while maintaining operations continuity.


Table 17.12 Mine Department Personnel

Description Number of Persons

Mine Operations Mine Superintendent 1 Mine Shift Foreman 4 Drill/Blast Engineer 1 Training Foreman 2 Shovel/FEL Operators 5 Truck Drivers 29 Rotary Drill Operators 5 Support Equipment 16 Blasters & Helpers 3 Trainees/Labourers 2 Subtotal: 68

Mine Maintenance Maintenance Superintendent 1 Maintenance Planner 1 Shop Foreman 4 Maintenance Clerk 1 Mechanics 15 Welders 4 Apprentices 4 Light Duty Mechanics 2 Tire Boy 2 Lube/Service & Fuel Truck 4 Labourers 2 Tool Crib/Warehouse 0 Subtotal: 40

Engineering/Geology Chief Engineer 1 Chief Geologist 1 Grade Control Geologist 1 Mine Planner 1 Technician/Surveyor 3 Clerk 0 Subtotal: 7 Total: 115


18.0 OTHER RELEVANT DATA AND INFORMATION

18.1 Tailings Facilities, Waste Rock Dumps and Site Water Management

Detail design of the tailings facilities and site water management are presented in Klohn Crippen's reports “Site Water Management Design Report” dated February 19, 2007 and “Tailings Facility Detail Design Report” dated April 5, 2007. Feasibility level geotechnical design of the waste rock dumps is presented in Klohn Crippen's report “Feasibility Design of Tailings Facility, Waste Dumps and Site Water Management” dated February 8, 2006.

The site investigations carried out in 1979/1980 were extensive and provided most of the required information for design purposes. Additional site investigation work was carried out in 2005 and 2006 to augment this information.

The plant, tailings facility and waste rock dumps will be located near the open pit in the upper basin of Ruby Creek. The main features of the feasibility design are provided in the following sections.

18.1.1 Tailings Characterization

The tailings produced by the mill will be relatively coarse with nominally 26% to 33% fines content. The Neutralization Potential Ratio (NPR) among all the ore types tested ranged from 7 to 42, with a median of 23. The tailings are, therefore, concluded to have a very low potential for producing acidic drainage that would liberate metals. This is consistent with the low sulphide sulphur content of the tailings (less than 0.03%). Additionally, the existing pilot plant tailings facility has remained in a neutral state for 35 years.

Testing of tailings supernatant water show that most elements meet the BC Water Quality Guidelines (BCWQG) for the protection of freshwater aquatic life. Aluminium and iron are the only elements that have median values exceeding BCWQG.

The tailings facility plan allows pond water to be discharged to the environment only during peak spring runoff. With a 20 times dilution rate, the median and maximum concentrations of all elements fall below the BCWQG. Nevertheless, the tailing facility is suitably sized to store all the tailings water during the life of the mine with no release to the environment, if required.

18.1.2 Tailings Facility

A compacted cyclone sand tailings dam will be constructed across Ruby Creek to form the tailings storage facility downstream of the plant. The tailings will be transported by slurry


pipeline at a rate of 21,000 to 23,000 tonnes of solids per day. Free water will be reclaimed from the pond by a floating pump-barge and returned to the plant. Water reclaim will be maximized to reduce the volume of stored water.

A seepage recovery dam will be constructed downstream of the tailings dam. It will also serve as a settling pond for any sediment transported by runoff and dam construction, and provide a monitoring station for water quality sampling.

At the end of operations, 94 Mm3 of tailings (excluding cyclone sand used for dam construction) and 30 millions tonnes PAG waste deposited in the facility will reach an elevation of 1394 m. The total volume of free water will be 7 Mm3, giving a final average tailings/water at an elevation of 1397 m.

The required dam crest elevation for the base case water balance is 1401.1 m. However, for facility layouts, an ultimate dam crest at 1418.2 m has been selected to accommodate a conservative water balance with incomplete diversion of the drainage catchments. The ultimate dam could also provide for 4 more years of mining reserves.

The tailings dam will be built from compacted cyclone sand dam constructed by the centreline method with a crest width of 30 m, and downstream slope of 3H:1V. The maximum dam height is 140 m and the final crest length is 1500 m. A system of finger drains will be installed at the base of the cyclone sand to keep the water level in the dam depressed.

A starter dam of compacted till will be constructed at an elevation of 1340 m to store 1.5 Mm3 of water for mill start up and to store the first year of tailings production. Subsequent dam raises will be constructed using cycloned tailings compacted in cells. The technique of "cell construction" is well established and has been used at Highland Valley Copper, Kennecott Utah Copper, Southern Peru Copper and in the oil sand industry.

All vegetation and soil beneath the dams and tailings basin will be stripped for site reclamation at closure. The exposed surface extending 500 m upstream of the tailings dam will be inspected for highly pervious zones (talus or colluvial materials) that could cause increased seepage or piping of tailings into the foundation. Any highly pervious zones will be excavated and replaced, or suitably covered by lower permeability compacted till soils.

Several localized pockets of loose soils were encountered at shallow depths of less than 5 m. Prior to dam construction, additional drilling investigations will be carried out to assess the aerial extent of these loose zones. If required, the loose zones can be easily excavated. The final ground surface beneath the dam will be compacted.


Tailings deposited from the crest of the dam will maintain a separation of 400 m to 1000 m between the dam and the water pond. This will limit seepage and keep water levels in the pervious cyclone sand dam very low. Seepage into the pervious basalt in the base of the valley will be restricted by the overlying cover of overburden soils, augmented over time by the 75 m depth of tailings. Analyses indicate 10 L/s to 33 L/s of seepage through the dam and foundation will occur under normal operating conditions. This is consistent with experience at the Gibraltar and Highland Valley Copper tailings facilities.

The static and dynamic stability of the tailings dam exceeds Factors of Safety recommended by the Canadian Dam Association.

18.1.3 Waste Rock Dumps

Geo-chemical testing indicates that most of the waste rock will not be acid-generating, with a median NPR in excess of 3.7. This conclusion is corroborated by the neutral paste pH of the rock samples (typical paste pH 6.8 to 9.1) and the low sulphide sulphur contents (median = 0.04%) of the waste rock.

Detailed evaluation of the deposit geology and NPR results suggests that select zones of the pit may product up to 30 Mt of potentially acid-generating (PAG) waste rock. PAG will be placed in the tailings impoundment so that it is permanently submerged within 1 year of placement.

Non acid-generating (NAG) waste rock and overburden will be deposited in small dumps located in valleys above the open pit and in a large main dump to the east of the open pit at the upper end of the tailings facility. The low-grade ore will be stockpiled and will be integrated with the larger waste dump which is nearest the plant.

Pro-active measures will be implemented during dump operations to minimize the mobilization of metals under neutral leaching conditions. Upslope diversion ditches will be constructed to convey runoff around the dumps and completed dumps will be immediately capped with overburden soils to minimize infiltration.

Final reclamation of the dumps will occur during the milling of low-grade material in the last four years of mill operation. The waste dumps will be re-sloped and terraced to safe long-term angles while meeting closure land use objectives. Dump surfaces will be capped with an overburden soil cover to shed water to a perimeter ditch that will route the water away from the dumps.

The foundations at the dumpsites are composed of competent glacial tills, coarse colluvium or bedrock. All sites are rated as having a low failure hazard according to the British Columbia Dump Stability Classification System. Calculated static factors of safety exceed 1.3 for operating


conditions and 1.5 for closure conditions. Dump displacements for a 475-year return period earthquake will be less than 1 m.

18.1.4 Site Water Management Plan

Two major diversions will route natural runoff around the open pit, waste dumps and tailings disposal facility. Both ditches will originate above the pit area and run generally parallel to Ruby Creek on the flanks of the surrounding hillsides, rejoining Ruby Creek downstream of the seepage recovery dam. The channels are sized to convey the 200-year return period peak flow with suitable freeboard to account for variations in hydraulic performance. Riprap armouring is provided on steep channel sections where peak flood velocities may be sufficiently high to cause erosion of the native till soils.

The diversion channels will be constructed on variable till, colluvium and talus. Where required, the base of the channels will be lined with compacted till and/or an 80 mil LLDPE to restrict leakage.

A net total of 86 Mm3 of undiverted water from the project site will be sent to the tailings facility over the 22 year operating life. The vast majority of the inflow will be retained in the voids of the tailings and submerged PAG waste rock (66%), and as free water in the tailings pond (8%). The remaining losses are evaporation (13%), groundwater seepage (10%) and controlled releases to Ruby Creek (3%).

A total of 264 Mm3 of pond water will be re-circulated to the mill over the 22 year operating life. For comparison, only 9.5 Mm3 will be impacted by the waste dumps in the tailings catchment. Hence, the tailings supernatant should govern the water quality in the tailings pond.

A dam freeboard allowance is provided for all dams to accommodate the 200-year return period 30-day storm plus an additional 3 m to allow for the routing of a 30-day PMF (assuming the diversions fail) through an emergency spillway.

A permanent main spillway channel will be constructed on the right abutment of the tailings dam to convey extreme floods from the tailings facility during operation and to convey the flows from the tailings basin at closure. A spillway is also provided in the left abutment of the seepage recovery dam. All spillway structures are designed for a peak flow of 35 m3/s produced by the routing of the Maximum Probable Flood (PMF) from the entire undiverted catchment of the mine site through the tailings facility.

During operation, an operating spillway will be excavated at the right abutment of the tailings dam to route the flood flows to the main spillway channel. The initial operating spillway will be


installed after completion of the tailings starter dam construction. For subsequent dam raises, operating spillways would be excavated only in an emergency when a trigger water level is reached (when the pond level reaches within 5 m of the dam crest). These spillways will be 5 m wide and excavated 3 m below the tailings dam crest so that a 1 m minimum dam freeboard is maintained above the peak PMF flood level. Riprap armouring for the spillways will be permanently stockpiled at the right abutment of the tailings dam for immediate use.

At closure, the permanent spillway channel will be extended into the tailings basin along the right abutment of the tailings dam. The base elevation of the channel will be set to provide a 700 m minimum separation between the tailings pond and dam crest under average flow conditions, and maintain a 3 m minimum dam freeboard above the peak PMF flood level.

18.1.5 Operating and Monitoring Controls

The tailings dam and seepage recovery dam will be monitored by piezometers and settlement pins to record the phreatic levels in the dams and foundation soils, and to measure the deformations of the structures.

Pond levels will be recorded monthly and used, in conjunction with pond filling curves, to plan the operation of the facility. The performance of the diversion channels and the operation of the spillways will be monitored and reported on a regular basis. The proposed monitoring and operational controls include:

• Establishment of flow gauging stations and survey of high water marks to confirm the flow capacity of the channels and the design relationships between precipitation and flood flows.

• Weekly inspections of all channels and spillways to visually assess performance and provide advance warning of excessive erosion or seepage through the base and banks of the channels.

• Daily inspections of the channels above the open pit which could represent a hazard if a breach of the channels diverts flow into the pit.

• Each spring before and during snow melt, the diversion structures will be inspected and cleared of snow and ice as required to ensure proper operation.

The emergency spillway will be constructed using stockpiled materials, if the pond rises to within 5 m of the dam crest. Records of pumping from the seepage recovery pond to the tailings pond will be kept to assess the seepage rate from the tailings facility.

Surface water sampling and flow quantities will be monitored at various points through the mine area to evaluate water quality and maintain a site wide water balance. The water quality and flows


will be evaluated to determine the allowable controlled discharge of tailings pond or seepage water to Ruby Creek, primarily during spring runoff.

Groundwater quality will be monitored at wells located downstream of the seepage recovery dam. This will be assessed in conjunction with Ruby Creek water quality to determine the need (if any) for installing pumpback wells.

18.2 Mine Closure Plan

A preliminary mine closure plan is presented in the Environmental Assessment Certificate (“EAC”) Application for the project. Adanac received the Project EAC on September 10, 2007. Details of the mine closure plan are summarized below.

18.2.1 Tailings Facility

The diversion channels will be decommissioned to re-direct flow back into the tailings facility. The seepage recovery dam will also be breached to prevent water storage. A 10 m wide riprap-lined closure spillway will be constructed in the left abutment of the tailings dam. The spillway channel will extend along the left abutment to the edge of the tailings pond. The invert elevation of this spillway will be set to maintain a 700 m minimum separation between the tailings pond and dam crest (under average flow conditions) and designed to safely pass the probably maximum flood (PMF) from the entire upstream catchment.

The tailings are not anticipated to be acid-generating and exposed tailings beach and dam surfaces will be suitably reclaimed to prevent erosion and meet post-closure land uses targets. The 3H:1V exterior slope of the dam is flatter than the 2.5 H:1V maximum slope adopted for reclamation of cyclone sand dams at other mine sites.

Natural runoff within the catchment basin will dilute and improve the water quality of the tailing pond supernatant, with a full 10 times dilution achieved in about 10 years. More rapid dilution could be achieved by pumping all or part of the pond water into the open pit at the end of mining.

The expected annual seepage through the tailings dam is ten times less than the annual volume of water discharged through the closure spillway and, therefore, will be subjected to substantial dilution.

18.2.2 Waste Dump/East Ruby

The East Ruby waste dump will be re-sloped and terraced to a safe long-term stable angle while meeting closure land use objectives. The dump will be reclaimed as soon as practical as each


portion of the dump becomes inactive. Final reclamation of the dump will occur in the final four years of milling low-grade ore.

To minimize infiltration of surface water, the top of the East Ruby Waste Dump will be graded and capped with glacial till to shed water to a perimeter surface water collection channel. The channels will also intercept runoff from the catchment slopes above the waste dump. The water will be routed to the open pit or the closed tailings facility.

Overburden soils from the dump will be stored for use in site reclamation of the waste dump and tailings dam, as required. The remaining portions of the waste dump will be suitably reclaimed. To reduce the effects of neutral leaching, the dump surface will be graded and capped with glacial till to shed water to outer edges of the dump.

18.2.3 Open Pit

The open pit will ultimately fill to become a small lake. Pit walls will be left in a safe and stable condition. No re-vegetation program is planned for the pit walls as they will be bare bedrock exposures as currently exist over much of the pit area. Surface runoff from above the pit will accumulate in the pit to a maximum elevation of about 1425 m, the approximate elevation of the outlet at the eastern rim. A channel leading from the eastern rim to the tailings facility will be constructed to ensure long-term stability and positive control of the outflow to the closed tailings facility.

18.2.4 Temporary Mine Closure

Temporary mine closure for periods longer than one year will require that measures, as outlined below, be implemented to mitigate the environmental impacts of the Project

Stockpiled low-grade material, which has a higher potential for metal leaching, will either be processed in the plant or capped with overburden soils. Water from the tailings pond will be pumped to the open pit to flood the pit walls. Any other materials deemed to be PAG or have high neutral leaching potential would also be submerged in the open pit.

Seepage water quality from the waste dumps will be monitored and, if required, measures taken to mitigate the potential impact on water quality in the tailings facility. These measures include pumping seepage to the open pit for storage and/or capping exposed dump surfaces to reduce infiltration.

An annual net surplus of about 3.7 Mm3 of water will report to the tailings facility during closure, resulting in a 2 m to 33 m pond water level rise without release. The height of the pond water


level rise depends on the size of the impoundment at the time of closure. For the first year of closure, sufficient water will be pumped to the open pit to maintain a constant pond level without discharge. If closure extends beyond 1 year, the closure spillway will be installed to route the PMF from the tailings facility. Constant year-round release of water through the spillway will then occur.

The diversion channels will be maintained during temporary closure periods. The diversion flows will mix with the spillway flows to achieve an acceptable water quality in Ruby Creek.

18.3 Environmental Considerations

The environmental component of this feasibility was conducted by Klohn Crippen. The following is an excerpt from the February 2006 Ruby Creek Molybdenum Project Executive Summary, Environmental Assessment Certificate Application:

A two-stage approach was used to determine the environmental and socioeconomic effects of the project. First, socioeconomic and environmental components were screened to rank their importance. Components were ranked on whether they could interact with the project as well as whether they are important to the Taku River Tlingit First Nation (TRTFN), regulators and other stakeholders. The results of this screening generate Valued Environmental Components (VECs) and Valued Socioeconomic Components (VSCs). These were defined as:

• Fish and fish habitat • Moose • Game birds

• Woodland Caribou • Grizzly bear • Receiving water quality

• Stone’s sheep • Hoary marmot

Fish and fish habitat assessments were based on the removal of fish habitat and fish populations as a result of the direct location of the footprint of the various project components, or the reduction in flows in the rivers downstream of the project.

The baseline work identified an isolated population of 200 - 400 Arctic grayling that will be moved as part of the environmental management for the fish habitat. The project will also provide a no-net-loss fish habitat compensation plan for the loss of habitat in the Ruby Creek drainage. The fish habitat compensation plan replaces lost spawning habitat on the lower parts of creeks affected by placer mining where this habitat is no longer available to fish.

To protect the animals in the area, there will be a no-hunting policy for employees accommodated at the project site, and access to and within the mining lease will be restricted.


Woodland caribou habitat will be directly affected by project construction and indirectly affected by project noise. Personnel of operating mines in the Northwest Territories report that caribou become accustomed to mining activity and even use the mine roads to escape biting insects. The mine will disturb some calving habitat; however, it is felt that this will not limit the stability of the caribou population.

There may be some direct removal of Stone's sheep habitat, but regionally this is a small amount relative to the total potential range of the sheep.

Moose generally use the lower section of Ruby Creek, with some use of the wetland in the lower part of the project footprint. In general, high quality moose habitat is not disturbed by the project. Moose also become used to project activities and experience at other mines shows that moose migrate into the mine area no-hunting zone during the hunting season.

Grizzly bear ranges are extremely large and land use is subject to disturbance by human activity. The additional work being carried out in the upper Ruby Creek drainage affects only a small part of the grizzly bear habitat.

A hoary marmot colony will be moved to accommodate the tailings disposal facility. As marmot colonies have been successfully moved in other areas, the long-term effect on their population, which ranges from Alaska into the continental USA, is expected to be minimal.

Game bird habitat will be affected by the project footprint, though the birds' range and the current densities are not felt to be limiting on their populations. Therefore, the project's effect on game birds will be the displacement of some birds to adjacent habitat areas.

Provincial regulations provide water quality guidelines for the protection of aquatic life. The project operations plan is to discharge water from the tailings disposal facility during the freshet when the receiving environment has excess water. The receiving water quality will reflect the background conditions and will not be affected where Ruby Creek enters Surprise Lake. At closure, water from the project site may have concentrations of some metals at levels that are similar to the elevated background concentrations.

The Environmental Assessment Application was accepted for review on August 01, 2006 and was received from the Province of British Columbia on September 10, 2007.


18.4 Capital Costs

2007 Capital Cost Estimate Update

The total initial capital cost for the development of the Ruby Creek Project is estimated to be CDN$640 million. The cost estimate has been carried out to an accuracy of +15%. The estimate was prepared in house by professional engineers.

Adanac realized the importance of updating the capital costs to reflect current market conditions to support the updated feasibility study due to the escalation in project costs over the past few years. The detailed estimate produced for the 2006 feasibility study was used as a basis and updated with current labour rates and construction materials costs ie. electrical cable, concrete and steel. Quantities were updated where detailed engineering supported refining the quantities. Further detail was available from additional geotechnical work and civil design and this was included in the updated estimate. Approximately CAN$130 million has been committed to date for process, mining and mobile equipment, and these costs have been included in the updated capital cost estimate. Committing to the process and mining equipment at an early stage of the project has eliminated the highest risk component in current estimates due to the volatility of these costs.

The resulting capital cost is a true reflection of current costs.

A summary of the major costs is shown in Table 18.1. Detailed descriptions of capital costs are in the appropriate reports.

Table 18.1 Capital Cost Summary

Cost Centre Total Construction Cost (CDN$)

Atlin Infrastructure

Property Purchase 184,069 Fencing / Infrastructure 100,000 Secure Storage (2 Containers) 8,000 Atlin Office 500,000 Atlin Permanent Housing Included in Owners costs 0

Total Atlin Infrastructure: 792,069

Ruby Creek Site Infrastructure

Temporary Roads and Grading 1,000,000


Construction Camp Camp Transport / Mobilization 7,750,000 Camp Rental 6,927,000 Maintenance Building 300,000 Catering and Consumables 13,671,000 Fuel and Propane 8,937,500 Water Supply 100,000 Sanitary Sewer 50,000 Solid Waste Disposal 108,500 Fuel Storage 220,003 Temporary Communications 310,000 Fisheries Compensation 400,000 Wildlife Management (2008) 120,000 Total Construction Camp 38,894,003 Permanent Facilities Construction Aggregate Supply and Material Processing 10,825,000 Batch Plant 1,900,000 Access Road 3,100,000 Bridge 500,000 Water Supply Bulk Earthworks Contractor Mob/Demob 2,000,000 Seepage Recovery Dam 12,000,000 Diversion Ditches and Rip Rap Production 14,836,000 Employee Village Access Road 2,500,000 Tailing and Roads and Sediment Ponds 24,950,959 Plantsite Civil and Piles 27,342,349 Concrete 31,853,901 Structural & Arch 23,035,562 Mechanical and Piping 55,641,467 Electrical and Instrumentation 30,745,974 Permanent Camp 14,270,161 Indirects 60,489,088 Total Permanent Facilities: 316,270,422 Engineering and Construction Management ADANAC Consultants 600,000 All North Road Design and CM 1,224,859 AMEC Engineering and Procurement 19,956,666 Golder Mine Engineering 205,674 Klohn Crippen Tailings Dam/Environmental/Geotechnical 4,584,963

Construction Management 14,925,000 Vendor Representatives 3,250,000 Expediting and Inspection 3,000,000 Ledcor Pre Construction Planning 1,001,629 Total Engineering and Construction: 48,748,790


Procurement Gyratory Crusher 4,317,234 Apron Feeders 862,500 Rock Breaker 326,633 Overhead Cranes 1,443,885 Mobile Crane 500,000 HPGR 15,266,782 Cone Crusher 3,214,610 Cyclopacs 295,809 Ball Mill 19,043,104 Liner Handler 1,078,000 Regrind Mill 2,880,000 Flotation Cells 1,241,169 Samplers 88,000 Compressors 157,500 Agitators 84,800 Pressure Filter 488,000 Bagging System 683,900 Scrubber 222,300 Conveyors 4,750,000 Lime System 355,000 Dust Collectors 25,000 Reclaim Barge 2,231,550 Pumps 245,000 Air Handling 185,523 Thickener 402,500 Boilers 375,000 Standby Gensets 3,300,000 Structural Steel 24,000,000 Gensets 21,433,220 Total Procurement: 109,497,019 Total Construction: 515,202,323 Contingency 50,628,464 Risk 20,130,004 Total: 585,960,791 Mining Pre-Strip 24,039,207 Mining Fleet 25,000,002 Contingency 5,000,000 Total: 640,000,000

2006 Capital Cost Estimate

The capital cost estimate produced for the 2006 Feasibility Study was a high level estimate prepared in accordance with industry standards for Feasibility Studies, to an accuracy of +/_15%..


The basis of the capital cost estimate was engineering designs and quantity take offs including: mine plans and haul profiles, process metallurgical testwork, design criteria, flowsheets and mass balance, equipment specifications, layouts and general arrangements, simplified P&ID’s, Single line diagrams and electrical distribution design.

Owner supplied equipment costs were obtained from quotations by suppliers. Mining Equipment was selected based on the mine plan, haul profiles and load/haul simulation software. Process equipment selection was completed based on extensive metallurgical testwork, simulation and design.

Construction bulk material costing is developed from vendor quotations and bills of materials that are taken off from piping line lists, electrical cable lists, valve lists and structural, mechanical, piping, electrical and instrument red lines produced by the discipline enginers. Engineered sketches were prepared to fully define piping, structural steel and concrete building services. Due to the high level of design for the tailings and water management completed to support permitting efforts these components were estimated to a very high level of accuracy.

Built up, burdened labour rates were developed from input from local and regional contractors.

The 2006 capital cost estimate was a build up of over ten thousand line items of detail including units, manpower productivity, built up labour rates, bulk, tagged (owner supplied) equipment, construction equipment and subcontract costs and included indirect and contingencies applied by work package.

18.5 Operating Costs

2007 Operating Cost Estimate

Operating Costs for the 2007 Feasibility Study Update were developed from first principles taking in to account the increase in plant throughput from 900 tonnes per hour to 1100 tonnes per hour.

The mine operating costs were developed using enhanced mine plan, equipment selection and operating hours. Process operating costs could be defined to a very high level of accuracy as flow sheets and equipment selection is frozen allowing reagents and consumables to be accurately defined, especially with respect to diesel consumption, a precise estimate of electrical power developed and fuel consumption was prepared. Consumable costs were updated from budget quotations from Vendors.


Labour costs were developed for an updated staffing and operating plan using current burdened labour rates from current collective bargening agreements in Northern British Columbia and the Yukon.

G&A costs were updated using current labour, camp and transportation costs based on the increased level of detail available since the original feasibility study.

After the operating costs were updated from first principles they were benchmarked against similar operations as a check. Comparisons were within the accuracy of the estimate. The costs were reviewed by a reliable independent reviewer, Mr R. Pendreigh P. Eng.

Process operating supply costs are based on budgetary prices from vendors of the consumables and reagents. The costs for general and administration (G&A) includes mine management, transport, insurance, warehouse and security personnel and general management.

Tailings dam construction costs have been estimated from other published costs from mines using similar construction techniques.

The total operating cost for is estimated to be CDN$11.71 for the first four years of full production and CDN$7.90 per tonne of mill feed for the remaining years of operation. Commissioning and pre-stripping activities completed in Year 1 have been accounted for in the capital cost estimates. The operating costs for mining, processing, power, tailing dam operation, general administration and Adanac’s related cost were prepared by in house professional engineers and reviewed by independent qualified persons to confirm that the work conforms to good engineering practice. Tables 18.2 and 18.3 present the operating cost summary.


(First Five Years of Full Production)

Description Operating Cost (CDN$/t ore milled)




(After Five Years of Full Production to End of Life [Year 21])

Description Operating Cost (CDN$/ tones of ore milled)


2006 Operating Cost Estimate

Operating Costs for the 2006 Feasibility Study were completed from first principles and developed to a high level of accuracy. Operating costs for the mining fleet were based on equipment usage and defined from currently accepted equipment performance characteristics. Operating costs for the process and G&A were developed from detailed reagent and consumable usage as defined from the metallurgical testwork and based on quotations from suppliers. Operating costs for G&A were developed from first principles with manpower loading and current burdened labour.

The mine has been considered to be owner-operated. The average unit mining cost was determined to be $1.39 per tonne mined or $3.94 per tonne of ore milled for the first four years of full production and $2.47 for the remaining years of operation.

18.6 Market

CPM Group’s report, “The Sustainability of Recent Molybdenum Prices” dated October 22, 2007, forms the basis for the molybdenum market assessment which follows below.

Structural shifts in the supply and demand of molybdenum have revived the molybdenum market after nearly a decade of depressed fundamentals and prices. Many of the underlying themes that drove prices sharply higher in 2004 are still present in the market. As of middle of October 2007, the spot price for molybdenum stood close to US$30.50 per lb.


18.6.1 Supply Fundamentals

Several aspects of the molybdenum market make it a unique industry that is difficult to comprehend at first glance. Molybdenum originates from two distinct sources each of which adhere to different market fundamentals.

Historically roughly 60%, of the world’s molybdenum supplies has come from copper deposits with included molybdenum mineralization. In these cases, molybdenum is produced as a copper by-product. Copper producers can exploit high grade molybdenum mineralized zones in copper ore bodies or add molybdenum recovery circuits, if molybdenum prices are high enough to justify the expense. This type of selective mining is not sustainable throughout the life a mine, but it can lead to large swings in the available supply of molybdenum from year to year. As an added complication, by-product producer’s output is largely influenced by copper prices. This has added to the volatility in molybdenum production and price. In the future, by-product producers could have a somewhat reduced effect on the molybdenum supply and prices, as their market share is forecast to fall to an average 52% over the ten year projections prepared by CPM Group.

Primary molybdenum deposits compose the majority of the remaining sources of production. Primary molybdenum producers, in contrast to copper producers, have historically acted as swing producers because, in the past, operating costs per pound of molybdenum have been lower for by-product producers than for primary producers. Depending on the price of molybdenum, even large-scaled primary producers have brought their operations on and off line. This has shifted in recent years, as more recently, it is the smaller by-product and primary producers which are better characterized as swing producers due to higher operating costs. The increased number of primary producers responding to the higher demand for molybdenum has helped to stabilize prices.

Increased demand for molybdenum, notably over the past four years, has stimulated a substantial restructuring of the molybdenum industry. Operating costs for both primary and by-product producers have also risen as miners seek to develop deposits that were once not economically viable. Practically all molybdenum producers are now facing operating costs that are greater than can be supported by the long-term average price for molybdenum of US$3.80 per lb. Based on forecasts of ore grades mined and higher actual unit costs production costs, cash costs for molybdenum production are expected to rise in the direction of US$12.00 per lb for some primary and by-product producers. The growth in demand and the resulting increase in molybdenum, is expected to offset these rising costs, however.

18.6.2 Supply Outlook

Under CPM Group’s Base Case scenario over the next three years, a deficit in the molybdenum market is predicted. Years of low investment levels in exploration and development have


contributed to an increasing number of bottlenecks throughout the molybdenum mining industry, which has led to a lag in bringing new mines on stream. Reportedly, new equipment orders are facing lead times as long as three years, as opposed to the normal six-month delivery period. In addition, output from China has been a growing source of uncertainty in the global molybdenum market, as China uses more of its molybdenum itself.

The Chinese government has been tightening regulation and control over not only molybdenum production and trade, but other mineral resources as well. In late December 2006 the Chinese government issued a revised version of its “Catalogue of Projects for which Land Use if Prohibited.” Molybdenum mining and smelting projects fall on this list. Thus approvals of new operations or expansions at existing operations for molybdenum mining or smelter projects is prohibited, except for projects that are upgrading to meet government standards. Present policy makes it impossible, or at the least very difficult, for any sizable new mine or expansion to obtain the necessary mining permits to come online. Some cities may choose to ignore the national policy and issue new permits. However, these would be smaller projects that could slip under the government’s radar. China’s production of high quality finished steel products and its domestic consumption provide strong internal demand for molybdenum. Given recent policy changes, it is unreasonable for the market to expect excess output from Chinese producers to meet the shortfall of western world production.

Fresh output from primary producers, which are expected to play a large role in filling the gap between mine production and demand, will begin to trickle into the market beginning in 2009. As for by-product production, the continued use of the leach-solvent extraction-electrowinning (SX/EW) process will limit the potential molybdenum recovery from some new copper projects, as the SX/EW process does not produce any molybdenum as a by-product. Considering the demand projections outlined in CPM Group’s base case scenario, a shortfall in mine production is forecast to remain until 2011. A ramp up at several large-scaled operations and additional supplies from secondary output should move the market into a market surplus in 2011. A moderate supply surplus, averaging less than 20 million pounds per year or a little over 3% of globally demand, is forecast from 2011 through 2016.

At this preliminary stage of planning it is difficult to gauge whether or not some of the expected molybdenum projects, which are forecast to contribute to the supply surplus in the later years of the projection period, will be able to overcome the obstacles inhibiting these projects from coming on stream. Permitting, economic, technical, and political hurdles plague a few of the proposed projects. If these issues are not properly addressed, these projects could face multiple setbacks and cost overruns, or may even be slashed before the mine is ever brought to production.


18.6.3 Demand Fundamentals

Molybdenum is primarily used as an alloying agent in a wide range of steels and alloys. The grayish non-toxic metal is employed in various steels, including many stainless steel grades because of its durability, strength, and robust qualities. Molybdenum alloys are resistant to extremely high temperatures as molybdenum has both a very low thermal expansion coefficient and one of the highest melting points of all elements. These qualities in conjunction with molybdenum’s other properties limit consumers’ ability to substitute for other metals in its numerous applications. Demand has been not only growing in its principal end uses, but demand for molybdenum has been evolving as many industries have sought to develop new materials that benefit from its alloying properties.

18.6.4 Molybdenum in Steel

Steel is an alloy of iron and carbon. Other alloying elements such as chromium, nickel, manganese, tungsten, vanadium, cobalt, silicon and molybdenum can be utilized in order to meet desired properties. Steel types can be broadly classified into three categories: Carbon steels, low-alloy steels, and high-alloy steels. Within each of these categories there are different types of steels, which come in various grades. This report looks at individual steel types. These include stainless steel, full alloy steel, carbon steel, tool and high speed steel, and high strength low alloy (HSLA) steel. Demand for molybdenum by the steel industry totalled 293 million pounds, or 69.9% of total demand in 2006.

Stainless steel is an iron - carbon alloy with a minimum of 10.5% chromium content. Molybdenum is used as an alloying element in stainless steels to enhance heat resistance, strength, malleability, corrosion resistance and to achieve thinness. Production of stainless steel is projected in the base case scenario to grow at a CAGR of 5.5% though 2016, after growing at a CAGR of 6.0% from 1995 to 2006. Demand for molybdenum by the stainless steel industry is projected to grow at the same pace as stainless steel production. In 2006, stainless steel accounted for 27% of molybdenum demand.

Full alloy steels are difficult to classify. They include structural and alloy engineering steel, Hatfield steel, maraging steel, carburizing steel, rail steel, spring steel, nitriding steel, bearing steel, pressure vessel steel, casting steel, weathering steel, and high strength steels. The molybdenum content of full alloy steels varies widely and is dependent on the steel’s end use. Some room for substitution from molybdenum to other alloying elements does exist. Demand for molybdenum is expected to grow at a slower rate than global GDP and is projected to grow at a 3.9% CAGR through 2016.


Molybdenum is used in tool steels to provide resistance to cracking, hardness, toughness, machinability and corrosion resistance. The amount of molybdenum in tool steel by weight can vary from 3% to 8% by weight, with higher molybdenum content steels produced primarily in the United States, European Union and Japan. Demand for tool steel tends to follow global economic conditions. As such, demand for molybdenum in tool steel could grow roughly at the same pace projected for global GDP growth, 4.8% per annum.

HSLA steel is used by the transportation, construction and energy industries in a wide range of applications from oil pipelines to automotive engine supports and bridge construction. Pipeline construction growth is expected to remain robust throughout the projection period, growing at approximately 4.0% per annum. This growth is projected to be led by Russia, China and India, with significant expansions in the United States as well. Molybdenum use in HSLA steels is projected to grow at a CAGR of 4.6% as a result of strong expected growth in the primary end uses of HSLA steel.

In carbon steels, carbon is the primary alloying element, adding strength and hardness to iron. Molybdenum is not used intensively in carbon steel production. When it is used, it enhances strength and hardness in addition to providing heat and corrosion resistance. Demand for molybdenum in the carbon steel industry is expected to grow at a 5.0% per annum throughout the projection period. This reflects strong demand for carbon steel, led by the emerging markets of China and India.

18.6.5 Other Molybdenum Applications

Molybdenum is used substantially in the petroleum refining and plastic industries. It is used in various steps of the refining process, especially hydroprocessing – the removal of sulfur, nitrogen, and other impurities from crude oil with the aid of hydrogen. The results of these processes are energy products such as gasoline, jet fuel, diesel, kerosene, and others. The forecasted 7% CAGR through 2016 for molybdenum demand in the catalyst industry is driven by the following factors: the construction of new refineries required to meet increased demand for energy products by emerging economies, the increased use of heavier crudes, environmental regulations, and molybdenum’s use in the direct liquefaction process in transforming coal to liquid energy products. Molybdenum’s use in the catalyst sector is expected to be strong.

A superalloy is an alloy that has high temperature qualities, high creep strength, and oxidation resistance. The term was first coined after the Second World War to describe alloys used in super chargers and aircraft engines. Although superalloys are still dominantly used in their traditional markets, usage has expanded to gas turbines, chemical plants, and petroleum plants. Superalloys are estimated to comprise 50% of an aircraft engine’s weight. They are usually made to order and are primarily composed of nickel, an estimated 80% of superalloy content. CPM Group’s


forecast of 5.5% CAGR through 2016 is heavily influenced by activities in the nickel and airline industry. The aircraft manufacturing industry is expected to grow at 5.5% through 2025.

Molybdenum use as a pure metal or molybdenum-based alloy is dominated by its use in the lighting, coatings, and glass industries. It is often alloyed with nickel and chromium to produce highly resistant coatings. These coatings improve the wear and friction properties of automotive parts such as gears, synchronizers, and piston rings. In lighting, its oldest application as a metal, molybdenum’s elevated temperatures strength, creep resistance and chemical compatibility with glass provide it with a significant advantage over substitutes. Recent price increases have caused substitution to nickel and titanium in emerging applications such as liquid crystal displays. Molybdenum demand in this sector is expected to grow at not more than 3% per annum through 2016.

Cast iron, also known as an iron-carbon-silicon alloy, refers to a family of multi-component ferrous alloys containing primarily iron, carbon, silicon, as well as other major and minor alloying elements. Molybdenum increases cast iron’s strength, heat resistance, and creep resistance. Due to molybdenum’s sparing use in fabrication, its demand here is relatively inelastic. Molybdenum demand growth in this sector is expected to almost match the pace of global economic growth. Cast iron growth is forecasted at 4.6% through 2016, driven by growth in China, India, and Asia-Pacific.

Molybdenum as a disulfide is used in lubricants due to a lower coefficient of friction than other lubricants, durability to withstand high temperatures, high yield strength, and a strong affinity for metallic surfaces. The global forecast for lubricant demand growth is 2.3% per annum through 2016, dominated by growth in lubricant use as engine oils as a result of growing manufacturing and automotive markets in India, China, and Russia.

Molybdenum is used in molybdate base pigments because of its stable color formation and corrosion inhibition properties. Molybdate pigments could be bright colors based on molybdenum oranges prepared by co-precipitating lead chromate, lead molybdate, and lead sulfate. These light and heat-stable pigments have a wide range of colors from bright red-orange to red-yellow, and they are used in paints, inks, plastics, rubber products, and ceramics. The use of lead and the toxicity of molybdenum present challenges in this market segment. Molybdenum demand in the pigment industry is expected to grow at 3% CAGR through 2016.

Molybdenum’s use in chemicals and other applications is dominated by its use as a smoke suppressant in PVC cabling. Molybdenum stabilizes char in combustible situations and thus prevents the formation of smoke particles. PVC is made through the polymerization of a vinyl choride monomer. PVC is used in construction, housing, automobiles, airplanes, and medical


devices. Molybdenum demand growth is expected to be 4.0% – 4.8% through the projected period, closely following GDP growth and influenced by the housing and construction industries.

18.6.6 Substitutes

Possible substitutes for molybdenum include chromium, tantalum, vanadium, columbium (niobium), tungsten and boron. The metals most frequently used as molybdenum substitutes in steel are columbium and vanadium. Substitution away from molybdenum typically comes at an economic or performance loss in most situations. The last published price for Brazilian polychlore, the columbium ore mineral, is from 1981 and in 2006 realized costs per pound were around US$45 per pound. Approximately 82% of mine production is from Brazil. Substitution from molybdenum to vanadium typically results in a lower melting point and lower corrosion resistance. In many applications the impulse may be to substitute other metals with molybdenum, despite higher molybdenum prices, due to increases in the prices of these metals as well. Cases in which there is little room for substitution to molybdenum include the use of chromium in stainless steel or tantalum in capacitors.

18.6.7 End-Use Industry Analysis

Approximately 38% of molybdenum demand comes from the energy industry. Molybdenum is used extensively in nuclear power plants, coal to liquids facilities, and petroleum refining, pipelines, and drilling equipment. The end uses of molybdenum that relate to energy include high strength low alloy steel, stainless steel, carbon steel, tool steel, full alloy steel, and catalysts. Worldwide energy demand has grown at a 2.6% per annum rate between 2001 and 2006, double the 1.3% per annum growth rate in energy requirements in the previous five year period. Molybdenum demand grew at a CAGR of 7.1% per annum from 2001 to 2006, after growing at a CAGR of 3.7% per annum in the previous five year period. Demand for molybdenum by the energy industry is expected to remain robust, accounting for 38% to 40% of total demand throughout the projection period.

Residential and commercial construction each utilize molybdenum. In the United States growth in the construction industry has traditionally been led by residential construction, which has been trending lower since May 2006. This decline has been offset in part by growth in the commercial construction sector, however. The value of the construction industry exceeded US$4 trillion in 2004, with significantly higher costs in the European Union and the United States. Spending is the most accurate way to gauge the pace of construction growth globally. Engineering News reported costs of US$550.00 per square meter in New York compared to US$70.00 per square meter in Shanghai. China, India and other emerging markets with lower construction costs are projected to drive the construction industry and sustain demand for molybdenum in the construction industry.


Molybdenum’s use in the transportation industry is a testament to its versatility as a metal and its role in contemporary industrial technology. It is used in all modes of transportation whether by land, air, sea, or rail. In the railway sector, it is used in wheel seats, brake pads, and locomotives engines. Molybdenum’s use in the aerospace sector is due to its heat and creep resistance properties in constructing airplane engines. These engines are made with superalloys, of which molybdenum is an important component. Shipbuilders use molybdenum in duplex steels due to its corrosion resistance properties, enabling them to transport a wide variety of materials on sea. Finally in automotives, molybdenum is used in lighting caps, turbocharger housings, automotive engines, automobile bodies, and suspension springs. Growth of molybdenum use in the transportation industry is expected to almost match GDP growth at 4.6% per annum, resulting in 96.9 million pounds of molybdenum demand in 2016.

18.6.8 Price Outlook

CPM Group forecasts the price of molybdenum to remain strong over the next ten years, above previous historical averages, due to tight demand and supply conditions covered in CPM Group’s Base Case scenario. Prices are unlikely to revert to their levels in the 1990s as operating costs for both primary and by-product producers have risen sharply. Prices may reach US$34.00 per lb in 2008, due to the predicted shortage in supply. However, prices are expected to hold above US$28.00 per lb through 2010. By 2016 prices may fall further. This is demonstrated in Table 18.4.


Table 18.4 Molybdenum Price Outlook

Year Price (US$/lb Mo)

2009 32.25 2010 28.00 2011 23.00 2012 21.75 2013 19.50 2014 16.00 2015 15.00 2016 14.75

Given the potential for the molybdenum supply outlook to be bleaker than projected in CPM Group ’s main scenario, toward the end of the projection period, prices could continue to be supported at higher levels. This is presented graphically in Figure 18-1.

Figure 18-1 Real Molybdenum Prices and Word Supply and Demand Balance

220

300

380

460

540

620

700

780

1995 1998 2001 2004 2007p 2010p 2013p 2016p0

6

12

18

24

30

36

42Million Pounds $US/Lb.

World Demand (LHS)

World Supply (LHS)

Actual Projections

CPM Group Base Case: Real Molybdenum Prices and World Supply and Demand BalanceAnnual, Projected through 2016p

Molybdenum Prices (RHS)


18.7 Economic Model

CPM Group prepared an Economic Model for the Project based on the following assumptions for the base case:

• Project construction starts in 2007 with commissioning in early 2009;

• Production commences in the second quarter of 2009;

• A US$/CDN$ exchange rate based on a consensus view of the leading CDN$ market makers and their outlook for the US$/CDN$ exchange rate sourced from Bloomberg;

• A roasting charge of US$1.77/kg of Mo contained in concentrate to convert MoS2 to MoO3 and a conversion loss of 1.0%;

• Concentrate shipped to a roasting facility in North America (to be determined);

• Declining molybdenum metal price from US$32.25 per pound in 2009 to US$14.75 per pound in 2016 and remaining flat at that level thereafter in the base case;

• The model was prepared on a pre-tax basis.

18.7.1 Net Present Value and Internal Rate of Return Summary

CPM Group developed a spreadsheet that could be used by Adanac to evaluate the sensitivity of the Project to various inputs. The ability to vary these inputs allows the impact of updated field results as well as changes to capital or operating costs, metallurgical test work results, exchange rates or the outlook for molybdenum prices to be examined quickly and the effects on the Project determined.

Capital, initial or sustaining, was considered in the year that it was spent. The contingency reserve was based on firm purchase orders or quotes, or refreshed estimates from detailed engineering. Consequently the contingency amount varies from 5% to a maximum of 15%, depending on the confidence of the design and the design category.

Revenue was calculated using the determined grade and molybdenum price assumption, and adjusted for marketing, transportation costs and metallurgical recovery.

Molybdenum price assumptions were based on the price outlook developed by the report “The Sustainability of Current Molybdenum Prices” dated October 22, 2007 prepared by the Commodity Market Research Department at CPM Group. The base case outlook for molybdenum prices contained in that report are in the Table 18-4.


The metallurgical recovery rate of 90% was selected based on metallurgical test work completed for the Ruby Creek Molybdenum Feasibility, April 2006, by Wardrop Engineering Inc. The gross revenue was calculated based on the recovered molybdenum. The net revenue was calculated from gross revenue less roasting, transportation and marketing charges.

In the base case the Project has an IRR of 18.9% and an NPV of $295 million at an 8.00% discount rate. Payback is 3.2 years.

18.7.2 Sensitivity Analysis

Sensitivities to molybdenum prices, the US$/CDN$ exchange rate, capital and operating costs on the IRR and NPV were considered. The most significant variable influencing the economic model is molybdenum prices. In addition, specific economic sensitivities were also run for the conditions described below.

3-Year Historical Average Molybdenum Price – The average monthly molybdenum price over the past 3 years is US$28.46 per lb. Using this price flat throughout the life of the Project, yields an IRR of 30.3% and a NPV of CDN$1.014 billion at an 8.00% discount rate.

Low Case Molybdenum Prices – The low case outlook for molybdenum prices contained in the CPM Group report reflected lower growth in demand for molybdenum, primarily in the stainless steel and pigments industries. Lower molybdenum prices would curtail some early stage or small-scale, high cost molybdenum projects, categorized by CPM Group as “possible” project, from coming on stream. This stems from the fact that lower prices would likely lead to a reduced influx of fresh supply. The Project IRR decreased to 12.3% and the NPV decreased to CDN$121 million at an 8.00% discount rate

High Case Molybdenum Prices - The high case outlook for molybdenum prices contained in the CPM Group report reflected higher growth in demand for molybdenum in the stainless and HSLA steel industries, lubricants and catalysts. Higher prices would also spur some additional supply from projects, now deemed possible. However, the supply response is muted due to delays in bringing production on-stream. Under the high case molybdenum price scenario, Project IRR was 24.6% and the NPV increased to CDN$444 million at an 8.00% discount rate.

Mine Operations with Grid Power Starting Earlier (From Year 3 Onwards) - The base case considers a conversion from diesel-fired electricity generation at the Project site over to accessing grid power in Year 5. This sensitivity considered introducing grid power to the Project 2 years’ earlier than in the base case, in Year 3. The Project IRR increased to 19.5% and the NPV increases to CDN$325 million at an 8.00% discount rate, from the base case.


Increase in In-Situ Grade by 15% - A sensitivity was carried out to determine the effect of a potential 15% increase in ore head grade. The case for a potential increase in ore grade can be made based on detailed comparative drill hole/bulk sample studies in the “Feasibility Study of the Adanac Molybdenum Project” prepared by Kaiser Engineers for Kerr Addison Mines Limited, Report 71-1, dated January 1971. While this work is not NI 43-101 compliant, the results from processing an approximate 10,000 tonne bulk sample at a 100 TPD pilot plant on site indicated that realized grades were higher than those predicted by drilling results by as much as 20%. For the purpose of a sensitivity study only a grade of increase of 15% was assumed. With the increase in grade in this scenario, the Project IRR increases to 27.6% and the NPV increases to CDN$ 533 million at an 8.00% discount rate.

18.7.3 Summary of Results

A summary of the results of the sensitivity analyses is presented in Table 18.5.

Table 18.5 IRR and NPV Summary Results

Case Description IRR NPV @ 8%

(CDN$ millions)

Payback Period (Years)

Base Case 18.9% 295.0 3.2 Sensitivities Historical Average Mo Price (Last 3 years’) 30.3% 1,014.7 2.9 Low Case Mo Price Scenario 12.3% 120.7 5.9 High Case Mo Price Scenario 24.6% 444.1 2.6 Capital Cost +15% 14.8% 213.0 3.8 Capital Cost -15% 24.8% 377.2 2.6 Operating Cost +15% 15.2% 185.6 3.5 Operating Cost -15% 22.4% 404.5 2.9 Specific Economic Sensitivities Mine operation with hydroelectric power starting inj year 3 19.5% 325.0 2.6

Increase in in-situ grade by 15% 27.6% 533.4 2.4

Long Term Mo Price 2016 onward $ US / lb

NPV @ 8% (CDN$

millions) $10.75 156.7 $12.75 225.9 $16.75 364.3 $18.75 433.4


Long Term Mo Price 2016 Onward $ US / lb

NPV @ 6% (CDN

Millions) $10.75 222.9 $12.75 315.2 $14.75 407.5 $16.75 499.8 $18.75 592.1

Long Term Mo Price 2016 Onward $ US / lb

NPV @ 10% (CDN Millions)

$10.75 104.2 $12.75 156.6 $14.75 209.0 $16.75 261.4 $18.75 313.8

Given that molybdenum prices are quoted in US$ and hence Project revenue, and operating costs are largely denominated in CDN$, the US$/CDN$ exchange rate is important. Indeed recent strength of the CDN$ has benefited Adanac by lowering the cost of long-lead capital equipment procured to date for the Project.

Full details of the Economic Model are contained as Appendix A.


19.0 INTERPRETATIONS AND CONCLUSIONS

The Project contains a valuable molybdenum-bearing mineral resource that can be economically extracted using proven mining methods and processing technologies, at current labour, equipment and material costs and also based on the projected prices of molybdenum in the future.

20.0 RECOMMENDATIONS

On September 19, 2007 Adanac announced that it would proceed with the development and construction of the Project and that it is proceeding to arrange the equity and debt financing to build and operate the mine. Adanac has received an Environmental Assessment Certificate on September 10, 2007 from the province of British Columbia. This Feasibility Study reaffirms the economic viability and financial sustainability of the Project.

It is recommended that Adanac continue to develop the Project through detailed engineering and construction.


21.0 REFERENCES

Adams Metals Ltd., Molybdenum Market Study, July7, 2005.

Adanac Molydenum Corp, Ruby Creek Molybdenum Project Executive Summary, Environmental Assessment Certificate Application, February 2006.

BC Mining Research Ltd., Preliminary Evaluation of Ultra-Fine Grinding for the Ruby Creek Molybdenum Project, February 10, 2005.

Blower, S. (2005): Technical Report – Mineral Resource Estimate Ruby Creek Molybdenum Project, report dated April 11, 2005. (AMEC Americas Limited).

Chapman, Wood & Griswold Ltd., 1971. Feasibility Study Kerr Addison Mines Limited Adanac Project (Volume II of II – Engineering and Economic Detail).

G&T Metallurgical Services Ltd., An Assessment of Metallurgical Response – Ruby Creek Project, December 27, 2006.

Golder Associates Ltd., Pit Slope Stability Considerations for the Ruby Creek Project, Adanac Moly Corp, Atlin, BC, February 10, 2006.

Golder Associates Ltd., Ruby Creek Molybdenum Project Mining Feasibility Study, March, 2006.

Golder Associates Ltd., Technical Report — Mineral Resource Estimate, Ruby Creek Molybdenum Project, February 17, 2006.

International Molybdenum plc website: www.internationalmolybdenum.com

Janes, R.H. (1971): The Geology of the Ruby Creek Molybdenum Deposit; in Chapman, Wood and Griswold, Economic Feasibility Study, Volume VII, pp 1-14 (unpublished).

Klohn Crippen Consultants Ltd., Ruby Creek Project — Feasibility Design of Tailings Facility, Waste Dumps and Site Water Management, February 8, 2006.

Monger, J.W.H. (1975): Upper Paleozoic Rocks of the Atlin Terrane, Northwestern British Columbia and South Central Yukon; Geological Survey of Canada, Paper 74 – 47, pp 1-63.


Neil S. Seldon & Associates Ltd., Marketing and Commercial Input for the Ruby Creek Molybdenum Project, August 2005.

Palmer, P., (2006): Technical Report Mineral Resource Estimate Ruby Creek Molybdenum Project, report dated February 17, 2006, Golder Associates Ltd.

Pinsent, R.H. (1980): Diamond Drilling Report on the Adanac Property, Adera 1, 4-8, Hobo 8, 19-20, 47 and Key 27 Mineral Claims, Atlin Mining Division; British Columbia Ministry of Energy, Mines and Petroleum Resources, Assessment Report #8861, 3 pages plus appendices.

Pinsent, R.H. (2005). Diamond Drilling Report on the Adanac (Ruby Creek) Property, British Columbia Ministry of Energy, Mines and Petroleum Resources, Assessment Report, dated January 31, 2005.

Pinsent, R.H. and Christopher, P.A. (1995): Adanac (Ruby Creek) Molybdenum Deposit, Northwestern British Columbia; Canadian Institute of Mining and Metallurgy Special Volume No. 46, pp 712-717.

SGS-MinnovEX Technologies Inc., Flotation Testwork Report for Plant Design, January 2005.

SGS-MinnovEX Technologies Inc., QEM SCAN Investigation of ADA NA C Molybdenum Products 0510-AD-547, December 19, 2005.

SGS-MinnovEX Technologies Inc., Ruby Creek Semi-autogenous Grinding Circuit Design, June 10, 2005.

Sinclair, A.J. (2005) Quality Control of the 2004 Adanac Moly Corp. Drilling Program, Ruby Creek Deposit. Unpublished consultant’s report to Adanac Moly Corp.

Sinclair, W.D. (1995): Porphyry Mo (Low-F-type), in Selected British Columbia Mineral Deposit Profiles, Volume 1 - Metallics and Coal, Lefebvre, D.V. and Ray, G.E., Editors; British Columbia Ministry of Energy of Employment and Investment, Open File 1995-20, pp 93-96.

Sutherland Brown, A. (1970): Adera, in Geology, Exploration and Mining in British Columbia, 1969; British Columbia Ministry of Energy, Mines and Petroleum Resources, pp 29-35.

Tennant, S. (1979): Adanac Drill Programme; British Columbia Ministry of Energy, Mines and Petroleum Resources, Assessment Report #7727, 5 pages plus appendices.


Wardrop Engineering Inc., Ruby Creek Feasibility, Process and Infrastructure Design and Cost Estimate, March 2006.

White, W.H., Stewart, D.R. and Ganster, M.W. (1976): Adanac (Ruby Creek) in Porphyry Deposits of the Canadian Cordillera, Edited by A. Sutherland Brown; Canadian Institute of Mining and Metallurgy Special Volume No. 15, pp 476-483.


22.0 DATE AND SIGNATURE PAGE

The Technical Report titled “Ruby Creek Project, Feasibility Study Update” was prepared and signed by the following authors dated December ●, 2007

Original signed by:

Rick Alexander, P.Eng

APPENDIX A ECONOMIC MODEL

Ruby Creek Feasibility Study Update - A1 - December, 2007

APPENDIX B CERTIFICATE OF QUALIFICATION

• Rick Alexander Certificate

report on feasibility study update ruby … study update ruby creek project, ... table 11.2 drill...

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