Western Prospector Group Ltd and Emeelt Mines Llc

WESTERN PROSPECTOR GROUP LTD AND EMEELT MINES LLC

INDEPENDENT TECHNICAL REPORT ON THE RESULTS OF A PRELIMINARY ECONOMIC ASSESSMENT GURVANBULAG URANIUM DEPOSIT, MONGOLIA

Compiled by:

Malcolm Buck, P.Eng. Bruce C. Fielder, P.Eng. Marek Nowak, MASc., P.Eng. Eugene Puritch, P.Eng. Mani M. Verma, P.Eng. Jane Spooner, P.Geo.

EMEELT MINES LLC November 27, 2007 MONGOLIA

SUITE 900 - 390 BAY STREET, TORONTO ONTARIO, CANADA M5H 2Y2 Telephone (1) (416) 362-5135 Fax (1) (416) 362 5763

Table of Contents Page

1.0 SUMMARY ...... 1 1.1 HISTORY AND DESCRIPTION OF THE PROPERTY...... 1 1.2 GEOLOGY AND MINERAL RESOURCES...... 1 1.3 MINING...... 4 1.4 METALLURGY AND PROCESS DESIGN...... 5 1.5 INFRASTRUCTURE AND SITE FACILITIES ...... 6 1.6 ENVIRONMENTAL, SOCIO-ECONOMIC CONDITIONS AND PERMITTING...... 6 1.7 THE MARKET FOR URANIUM ...... 7 1.8 PROJECT SCHEDULE AND MANPOWER ...... 8 1.9 PRELIMINARY ECONOMIC ASSESSMENT...... 8 1.10 CONCLUSIONS AND RECOMMENDATIONS...... 10

2.0 INTRODUCTION AND TERMS OF REFERENCE ...... 12 2.1 ACKNOWLEDGEMENTS ...... 13 2.2 UNITS OF MEASURE AND ABBREVIATIONS...... 13

3.0 RELIANCE ON OTHER EXPERTS ...... 16

4.0 PROPERTY DESCRIPTION AND LOCATION ...... 17 4.1 ENVIRONMENTAL CONSIDERATIONS...... 21 4.2 PERMITS...... 21

5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY ...... 22

6.0 HISTORY ...... 24

7.0 GEOLOGICAL SETTING...... 25 7.1 REGIONAL GEOLOGY...... 25 7.1.1 Central Mongolia Fold Belt ...... 25 7.1.2 Argun-Mongolia province...... 26 7.1.3 Stratigraphic Setting...... 27

8.0 DEPOSIT TYPES ...... 28 8.1 URANIUM IN VOLCANIC ROCKS ...... 28

9.0 MINERALIZATION...... 29 9.1 BASEMENT ROCKS (LOWER STRUCTURAL STAGE)...... 29 9.2 MESOZOIC ROCKS (UPPER STRUCTURAL STAGE)...... 29 9.3 STRUCTURE ...... 31 9.4 HYDROTHERMAL ALTERATION...... 31

10.0 EXPLORATION...... 33 10.1 2005 EXPLORATION PROGRAM...... 33 10.2 2006 EXPLORATION PROGRAM...... 34

11.0 DRILLING ...... 35 11.1 LOCATION AND SURVEYING...... 36 11.1.1 Field Location...... 36 11.1.2 Surveying ...... 36 11.2 DOWNHOLE SURVEYS...... 38 11.3 CONSTRUCTION OF DRILLHOLE DATABASE FROM RUSSIAN DATA...... 39 11.3.1 Digitizing of Russian drillholes ...... 40 11.4 DRILLING RESULTS...... 41 11.4.1 Resource Block 1-06...... 41 11.4.2 Verifications in Gurvanbulag Central Zone...... 43 11.4.3 Exploration near Gurvanbulag Central Zone...... 44 11.5 CONCLUSIONS OF WNP 2005 DRILL PROGRAM ...... 44

12.0 SAMPLING METHOD AND APPROACH ...... 46 12.1 DATA ACQUISITION ON CORE...... 46 12.1.1 Core Handling...... 46 12.1.2 Depth and Recovery Measurements ...... 46 12.1.3 Quick Log and Radiometric Scan...... 48 12.1.4 Geotechnical Logging...... 48 12.1.5 Radiometric Scan Log...... 49 12.1.6 Core Logging...... 49 12.1.7 Specific Gravity Measurements...... 53 12.1.8 Photography of Core...... 54 12.1.9 Core Storage...... 54 12.2 DRILL CORE SAMPLING...... 54 12.3 SHIPMENT OF SAMPLES...... 55

13.0 SAMPLE PREPARATION, ANALYSES AND SECURITY...... 57 13.1 SAMPLE PREPARATION AND ANALYSES...... 57 13.1.1 Alex Stewart Laboratory, Ulaanbaatar ...... 57 13.1.2 XRF Analyses in Ulaanbaatar...... 57 13.1.3 Activation Laboratories, Canada...... 58 13.2 QUALITY ASSURANCE AND QUALITY CONTROL PROGRAMS...... 58 13.2.1 Duplicates ...... 58 13.2.2 Blanks and Standards - Selection...... 60 13.2.3 Blanks and Standards - Insertion ...... 63 13.2.4 Re-assaying Procedures...... 63 13.2.5 Re-Sampling Procedure...... 65

13.2.6 Record Keeping for Traceability ...... 65 13.2.7 Data Storage and Security...... 67 13.3 COMPILATION OF RUSSIAN ANALYTICAL DATA ...... 67

14.0 DATA VERIFICATION ...... 68 14.1 CONTROL OF ASSAY RESULTS ...... 68 14.2 ASSAY DATABASE...... 68

15.0 ADJACENT PROPERTIES ...... 70

16.0 MINERAL PROCESSING AND METALLUGICAL TESTING ...... 71 16.1 METALLURGICAL TESTWORK...... 71 16.1.1 Composite Preparation...... 71 16.1.2 Leach Tests...... 73 16.1.3 Settling and Flocculation Tests...... 74 16.1.4 Solvent Extraction Tests ...... 75 16.1.5 Uranyl Peroxide Precipitation...... 75 16.1.6 Tailings and Environmental Data ...... 76

17.0 MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES...... 78 17.1 MINERAL RESOURCE ESTIMATION...... 78 17.1.2 Resource Estimation Methodology...... 81 17.1.3 SRK Mineral Resource Estimate ...... 82 17.2 MINERAL RESERVE ESTIMATION ...... 83

18.0 OTHER RELEVANT DATA AND INFORMATION ...... 84 18.1 MINING...... 85 18.2 MINING SEQUENCE...... 89 18.3 DEVELOPMENT AND PRODUCTION SCHEDULE...... 90 18.4 WASTE ROCK DISPOSAL...... 90 18.5 MINE SUPPORT FACILITIES...... 91 18.6 PASTE BACKFILL...... 91 18.6.1 Backfill Plant...... 91 18.6.2 Backfill Distribution ...... 91 18.7 FLOWSHEET DEVELOPMENT...... 91 18.8 SITE SELECTION FOR PROCESSING PLANT ...... 92 18.8.1 Approach to Tailings Disposal...... 95 18.9 INFRASTRUCTURE ...... 95 18.9.1 Road and Power Line...... 96 18.9.2 Site Services...... 97 18.9.3 Materials Shipment ...... 98 18.10 ENVIRONMENTAL, SOCIO-ECONOMIC CONDITIONS AND PERMITTING...... 98 18.10.1 Environmental Baseline Study...... 99 18.10.2 Previously Disturbed Areas ...... 99 18.10.3 Dewatering...... 100

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18.10.4 Groundwater ...... 101 18.10.5 Socio-economic Considerations...... 101 18.10.6 Safety, Health and Environment...... 102 18.11 THE MARKET FOR URANIUM ...... 103 18.12 CAPITAL COST ESTIMATE...... 105 18.13 OPERATING COST ESTIMATE ...... 105 18.14 PROJECT SCHEDULE AND MANPOWER ...... 106 18.15 ECONOMIC EVALUATION...... 106

19.0 INTERPRETATION AND CONCLUSIONS ...... 111

20.0 RECOMMENDATIONS...... 112

21.0 SIGNATURE PAGE...... 113

22.0 REFERENCES...... 114

List of Tables Page Table 1.1 SRK Classified Mineral Resources for the Gurvanbulag Central Zone, Saddle Hills Project, Mongolia, November 3, 2006 ...... 4 Table 1.2 Summary of Results of Preliminary Economic Assessment of the Gurvanbulag Deposit ...... 9 Table 1.3 Results of Sensitivity Analysis ...... 10 Table 2.1 List of Abbreviations ...... 13 Table 4.1 Saddle Hills Area Exploration Licenses ...... 18 Table 4.2 Details of Exploration Licenses ...... 21 Table 5.1 Average Annual Precipitation at Bayandun Weather Station...... 22 Table 9.1 Stratigraphic Succession in the Saddle Hills Area...... 30 Table 11.1 Summary Statistics of 2005 Saddle Hills Drilling Program ...... 35 Table 11.2 Total Downhole Deviations in 2005 Drillholes ...... 38 Table 11.3 WNP Drilling Results from Russian Resource Block 1-06 ...... 41 Table 11.4 Comparison of Results from WNP and Russian Drillholes...... 43 Table 11.5 Results of Exploration Drillholes in the Gurvanbulag Area ...... 44 Table 12.1 Core Recoveries for 2005 Drill Program ...... 47 Table 12.2 Lithology Codes for the Gurvanbulag 2005 Drill Program ...... 50 Table 12.3 Geology, Structural and Alteration Codes for 2005 Drill Logs...... 52 Table 13.1 Selection of Blank and Standard Samples ...... 60 Table 13.2 Results of New Assays on Control Samples...... 61 Table 13.3 Statistics for Standard Samples During First Round Robin...... 62 Table 13.4 Statistics for Standard Samples During Second Round Robin ...... 62 Table 13.5 Combined Statistics for the Standard Samples ...... 63 Table 16.1 Test Composites – Detailed Analyses...... 72 Table 16.2 Summary of Leach Test Conditions and Results ...... 73 Table 16.3 Summary of Uranyl Peroxide Precipitate Quality on a Percent Uranium Basis...... 76 Table 16.4 Treated Effluent Analysis and Monthly Arithmetic Mean Concentration Discharge Limits ...... 77 Table 17.1 Comparison of Results of Five Estimation Methods ...... 81 Table 17.2 Description of Block Models (Wrinkled and Unwrinkled) for Model 2 ...... 82

Table 17.3 SRK Classified Mineral Resources for the Gurvanbulag Central Zone, Saddle Hills Project, Mongolia, November 3, 2006 ...... 83 Table 18.1 Development and Production Schedule for the Gurvanbulag Project ...... 90 Table 18.2 Summary of Key Process Design Parameters...... 92 Table 18.3 World Uranium Requirements, 2005-2030...... 104 Table 18.4 World Natural Uranium Production by Country, 2005...... 104 Table 18.5 Long-term Price Projection...... 105 Table 18.6 Summary of Capital Expenditure...... 105 Table 18.7 Summary of Unit Operating Costs...... 106 Table 18.8 Summary of Results of Preliminary Economic Assessment of the Gurvanbulag Deposit ...... 108 Table 18.9 Gurvanbulag Preliminary Economic Analysis...... 109 Table 18.10 Results of Sensitivity Analysis ...... 110

List of Figures Page Figure 1.1 General Location Map...... 2 Figure 1.2 Location of Exploration Licenses ...... 3 Figure 4.1 General Location Map...... 17 Figure 4.2 Location of Exploration Licenses ...... 20 Figure 7.1 Geological Fold Belts in Mongolia...... 26 Figure 7.2 Uranium Provinces and Districts of Mongolia ...... 27 Figure 11.1 Locations of Drill Holes...... 37

Figure 17.1 Locations of Drill Holes and Area 1-06 C1...... 79 Figure 18.1 Underground Development and Shaft Locations...... 86 Figure 18.2 Plan of 260 Metre Level...... 87 Figure 18.3 Plan of 260 Metre Level Showing Estimated Volumes of Mineralized Material...... 87 Figure 18.4 Condition of Underground Equipment Following Dewatering Program...... 88 Figure 18.5 View of 9900N Crosscut on 260 m Level...... 89 Figure 18.6 Simplified Processing Flowsheet ...... 93 Figure 18.7 Alternative Locations of Processing Plant and Tailings Management Facility ...... 94 Figure 18.8 General Site Plan...... 97 Figure 18.9 Potential Project Development Schedule...... 107

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1.0 SUMMARY

The Gurvanbulag uranium deposit, located in northeastern Mongolia, lies within the Saddle Hills property, a group of exploration licenses held by Emeelt Mines LLC (Emeelt Mines) and Western Prospector Mongolia LLC. Both Emeelt Mines and Western Prospector Mongolia are wholly-owned by Western Prospector Group Ltd. (Western Prospector), a Canadian company based in Vancouver, British Columbia. The property is operated by Emeelt Mines.

Micon International Limited (Micon) has been retained by Western Prospector to prepare a preliminary economic assessment of the Gurvanbulag property as part of its plans to bring the Gurvanbulag uranium deposit to an underground production scenario.

The following report presents the results of the preliminary economic assessment in accordance with the requirements of Canadian National Instrument 43-101. The preliminary economic assessment is contained within the report prepared by Micon on behalf of Western Prospector and Emeelt Mines titled, Preliminary Economic Assessment, Gurvanbulag Uranium Deposit, Mongolia, dated September, 2007. The effective date of the present Independent Technical Report is November 27, 2007.

1.1 HISTORY AND DESCRIPTION OF THE PROPERTY

The Gurvanbulag deposit was explored and partially developed in the 1970s and 1980s by the uranium exploration arm of the Ministry of Geology of the former Soviet Union. The partially developed mine was abandoned in the early 1990s following the collapse of the former Soviet Union. The shafts were capped, surface infrastructure was removed or destroyed and the partially developed underground workings were allowed to flood.

Western Prospector/Emeelt Mines started to acquire exploration licenses and began exploration work in the Saddle Hills area in 2004. See Figure 1.1 for the general location of the property in northeastern Mongolia. Dewatering of the underground workings was undertaken in the second half of 2006. Eleven exploration licenses are held by Western Prospector and Emeelt Mines and a twelfth (3367X) is held by Adamas Mining LLC and in which Western Prospector LLC is earning a 70% joint venture interest. The location of the exploration licenses is shown in Figure 1.2.

1.2 GEOLOGY AND MINERAL RESOURCES

The Gurvanbulag uranium deposit is a shallow-dipping, tabular deposit with strike and dip extents of more than 2.5 by 2 km, respectively. The mineralized horizon consists of two distinct domains adjacent to the hanging wall and footwall contacts of a barren, obsidian- bearing horizon, within a dominantly felsic volcanic sequence. The mineralization appears to be predominantly stratigraphically controlled; however, vein-hosted mineralization is known to occur above and below the principal mineralized horizon. The Saddle Hills district is distinctive in the presence of the laterally extensive volcanic obsidian horizon below

rhyolites at Gurvanbulag and the laterally extensive uranium mineralization that is conformable to bedding in parts of this horizon.

Figure 1.1 General Location Map

Figure 1.2 Location of Exploration Licenses

The estimate of mineral resources for the Central zone of the Gurvanbulag uranium deposit on which the preliminary economic assessment is based was prepared by SRK Consulting (Canada) Inc. (SRK) and dated November 17, 2006, on behalf of Western Prospector, and was filed on SEDAR in June, 2007. The primary objective of SRK’s report was to prepare an independent estimate of uranium resources that is compliant with the Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Definition Standards on Mineral Resources and Mineral Reserves. The SRK estimate was conducted on a dataset that combined both Russian data and new data collected by Western Prospector/Emeelt Mines in the 2005-2006 drill hole program; the drill data included holes drilled by Western Prospector/Emeelt Mines to mid-March, 2006.

SRK reports mineral resources for the Gurvanbulag deposit at a cut-off grade of 0.07% U3O8, based on a long term uranium price of $47 per pound U3O8 and its own internal estimate of potential operating costs for underground mining. The classified resource estimate is shown in Table 1.1.

Table 1.1 SRK Classified Mineral Resources for the Gurvanbulag Central Zone, Saddle Hills Project, Mongolia, November 3, 2006

Hanging Wall Footwall Totals Thousand U3O8 Thousand U3O8 Thousand U3O8 Thousand Tonnes (%) Tonnes (%) Tonnes (%) pounds U3O8 Indicated Mineral 570 0.19 2,260 0.22 2,830 0.22 13,633 Resources Inferred Mineral 700 0.13 1,970 0.15 2,670 0.15 8,642 Resources

SRK (2006) describes the classification of the uranium resources as follows:

“The current Mineral Resources have been classified as Indicated and Inferred Resources. Indicated Mineral Resources include those estimated blocks with: (i) a minimum of 5 samples from at least two drillholes, used for the estimation, (ii) a maximum average distance between samples of 35 metres, (iii) a >50% chance of being above the cut-off grade, as defined by a kriged indicator model with a grade threshold of 0.07% U3O8, and (iv) falling within a manually digitized domain designed to contain closely spaced mineralized areas in the vicinity of the existing underground workings. All other estimated blocks were assigned to Inferred Mineral Resources.

“The largest uncertainty affecting the Mineral Resources is the lack of supporting documentation for the Russian data. On a metre-by-metre basis, the Russian data accounts for more than 92% of the data used in the resource estimate. While there is a strong correlation between new surface drill sampling by WNP [Western Prospector/Emeelt Mines] and the historical Russian data, there are still some indications that the Russian data are slightly biased on the high side, in places. This uncertainty, combined with the lack of supporting information for the Russian data, is the principal reason why a ‘Measured’ classification cannot be assigned to the Gurvanbulag resource.”

1.3 MINING

Work undertaken by Western Prospector/Emeelt Mines in 2006 included the dewatering of the underground workings at Gurvanbulag. On completion of the dewatering program, surface and underground examination shows that existing mine development comprises:

• Two 4-m diameter concrete lined shafts, approximately 1 km apart, and both to a depth of 260 m below surface.

• One 6-m diameter concrete lined shaft to a depth of approximately 285 m. This shaft is equipped with a manway and steel piping for dewatering, compressed air and water lines.

• Extensive lateral development on the 260 m level consisting of:

• A main and north haulage crosscut from the 6-m diameter shaft.

• Hanging wall haulage drift along 800 m of the mineralized strike length. • Mineralized zones access crosscuts on approximately 50-m centres. • Sill in the mineralized zones. • Footwall ventilation drift connecting the two 4-m diameter shafts. • Boxholes and drawpoints with raises to mineralized zones above.

In general, the shafts and lateral development was found to be in good condition and that little deterioration had taken place since the workings were allowed to flood in the early- 1990s.

For the purpose of this preliminary economic assessment, due to the shallow dip and relatively thin, tabular nature of the deposit, it is proposed that the inclined room and pillar mining method will be applied in the majority of the deposit. In-stope mining will utilize conventional, non-mechanized mining with handheld jacklegs, longtoms and stopers for drilling and slushers for mucking. Mineralized material from ore passes will be removed by load haul dump (LHD) units which will load rail cars. The rail cars will transport mineralized material to dumps near the main shaft. In some areas, small shrinkage or longhole stopes will be used to mine near-vertical lenses of mineralized material.

The configuration of the mineralized zones requires the use of flexible and selective mining methods, which can accommodate variability in width and dip of mineralization. Stopes will initially be accessed from existing development on the 260 m level. As mining proceeds, subsequent intermediate access levels will be developed from the internal ramp upward towards the 80 m level. Later in the mine life, the lower mining horizons will be accessed from the 440 m haulage level and intermediate access levels will be developed from the internal ramp, up to the 260 m level.

The mining plan envisages that mined stopes be backfilled in order to minimize ventilation air losses and, also, to minimize the footprint of the tailings disposal facility on surface. Backfilling will also help improve the overall stability, though it is not critical nor, on the basis of available data, is it considered integral, to maintaining stability.

Given the mineralization geometry and presently estimated mineral resources, a mine production rate of 1,500 t/d was selected resulting in a mine life of approximately 10 years.

1.4 METALLURGY AND PROCESS DESIGN

Metallurgical testwork for the Gurvanbulag project was carried out under the direction of Melis Engineering Ltd. (Melis) at SGS Lakefield Research Limited (SGS Lakefield) on samples of drill core received at the laboratory in June, 2006. The testwork program included grindability, leaching, liquid/solid separation, solvent extraction, precipitation and tailings preparation.

Melis also developed a simplified processing flowsheet for the purpose of the preliminary economic assessment.

1.5 INFRASTRUCTURE AND SITE FACILITIES

Golder Associates Ltd. (Golder) was retained by Western Prospector/Emeelt Mines to provide a preliminary assessment of potential sites for the location of a tailings management facility (TMF) at Gurvanbulag. Western Prospector/Emeelt Mines have provisionally selected Option 1 as the preferred location for the tailings management facility. It has been assumed that approximately 50% of the total tailings will be returned to the underground mine as paste backfill.

It is proposed that a new road will be constructed to connect the Gurvanbulag site with Choibalsan. The road, approximately 117 km long, is anticipated to follow the route of the proposed power line that will be constructed in order to connect the Gurvanbulag site with the power generating station at Choibalsan. The feasibility study and route design have been completed by the civil engineering firm, Avarga Zam Co. Ltd. of Mongolia.

Permanent power supply will be provided by a new 110-kV line from the power station operated by Dornod Energy System LLC (DES) at Choibalsan.

Accommodation is presently provided in a new facility that has been constructed 5.5 km south of the mine site. The facilities comprise a 200-person capacity camp that was constructed in mid-2006 by Geomandal Star LLC. Accommodation is provided in 57 gers, all of which are fully winterized and equipped with underfloor heating. This facility will be expanded in the operating phase of the project.

Emeelt Mines will acquire warehousing and storage areas in Choibalsan in order to undertake transhipment of bulk and other materials between rail and truck. For incoming materials, it is anticipated that shipments will be consolidated into containers for ease of handling and security. Truck haulage from Choibalsan to Gurvanbulag will be undertaken by contractors. It is proposed that product from the processing plant will be in sealed 205-L steel vessels and shipped from site in secured ocean containers.

1.6 ENVIRONMENTAL, SOCIO-ECONOMIC CONDITIONS AND PERMITTING

Mineral exploration and development of a uranium mining project in Mongolia are permitted within the framework of the following three principal areas of legislation:

• Minerals Law of Mongolia, 2006. • Law of Mongolia on Radiation Protection and Safety, 2001. • Law of Environmental Impact Assessment, 1998.

An environmental baseline study has been prepared by EcoTrade LLC in order to provide data on the geography, geology, hydrology, hydrogeology, soil, climate, flora and fauna, and the socio-economic conditions of the area. Field studies for this report were carried out in

2005 and additional monitoring continued through 2006 and 2007. Reports on the environmental impact assessment and an environmental protection plan and monitoring program for the power line from Choibalsan have been prepared by Ecos LLC.

A Community Relations Officer was engaged in August, 2006 with responsibility for liaison with communities and official representatives.

Western Prospector and Emeelt Mines have stated their commitment to conforming to Safety, Health and Environment (SHE) standards applicable in Mongolia or internationally, whichever is the higher. A full-time SHE technician is employed to provide input to the environmental sampling program, to liaise with government agencies and with consultants working on the project, to audit and advise on SHE matters and to train employees. In July through November, 2006, two individuals were retained through a consulting firm to assist with development of radiation monitoring, inspection and employee training programs and to provide a full-time presence on-site for program management, monitoring and training. A staff radiation expert was retained directly by Emeelt Mines in November, 2006.

The objectives for the site at closure will be the following:

• Minimal level of ongoing emissions of radiation above background levels. • Removal of all structures to ground level. • Installation of a suitable cover for the tailings disposal areas. • Underground mine workings to be allowed to flood. • Recontouring of the site to blend with the surrounding terrain.

1.7 THE MARKET FOR URANIUM

Mined uranium is used principally as the fuel for nuclear-powered electricity generating stations. Relatively minor non-electricity uses include small nuclear reactors in research, for nuclear-powered marine vessels, in space travel and for desalination; and in radioisotopes for medical research, industry and in smoke detectors.

The saleable product resulting from the mining and processing of uranium-bearing minerals is uranium concentrate, known as yellowcake. Yellowcake contains between 70 and 90 per cent by weight of uranium oxides and is also referred to simply as U3O8. Aggregate annual worldwide uranium requirements are estimated to be in the range of 170-175 Mlb U3O8 with the United States needing the largest portion (50-55 Mlb per year U3O8). Globally, there are more than 60 nuclear fuel buying companies and/or organizations distributed throughout North America, Europe, Asia/Pacific and South America, located in countries which are signatories to the non-proliferation treaty and, thus, eligible to purchase natural uranium concentrates produced within the Republic of Mongolia.

1.8 PROJECT SCHEDULE AND MANPOWER

It is anticipated that the majority of work carried out in year -3 will consist of exploration, detailed engineering and procurement of equipment with long lead time. A full feasibility study is also planned to be completed during year -3, as is work related to securing the operating permit. Construction leading to production will be undertaken in years -2 and -1.

During the production period, it is estimated that a total of 531 people will be employed at the mine, 24 of whom will be expatriates and 507 Mongolian nationals. During the earlier years of the operation, a larger number of expatriates will be required and these will also provide training for National individuals in specific skills. It is anticipated that the company will be able to operate within a 90:10 constraint for National to expatriate employees.

1.9 PRELIMINARY ECONOMIC ASSESSMENT

At the request of Western Prospector, Micon has prepared a preliminary economic assessment of the Gurvanbulag deposit that is based on the mineral resources estimated by SRK and reported in its Independent Technical Report and Resource Estimate for the Gurvanbulag Deposit; Saddle Hills Uranium Project, Mongolia, dated November 17, 2006.

The preliminary economic assessment is based on indicated and inferred mineral resources estimated by SRK for the Central Zone of the Gurvanbulag uranium deposit, on the results of metallurgical testwork completed in December, 2006 and on the results of the dewatering of the underground workings at the end of 2006.

The results of the preliminary economic assessment are considered by Western Prospector to be a material fact with respect to the company. The preliminary assessment is preliminary in nature; it includes inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves; there is no certainty that the preliminary assessment will be realized.

The preliminary economic assessment is based on a plan to mine and process an average of 1,500 t/d of mineralized material that results in mining all of the presently estimated 5.4 Mt of resources over a project life of about 10 years.

The estimates of capital and operating costs are combined in the discounted cash flow evaluation for which a summary of the base case is shown in Table 1.2. Capital and operating costs have been estimated to a level of detail appropriate for a preliminary economic assessment, in this case to an overall level of +/-25%.

The economic evaluation is treated on a full project basis, i.e., assuming 100% equity financing and, for the base case, a uranium price of $185/kg U3O8 (equivalent to $84/lb U3O8) has been assumed.

Table 1.2 Summary of Results of Preliminary Economic Assessment of the Gurvanbulag Deposit

Item Unit Value Pre-production capital cost $M 228.7 Life of mine capital cost $M 269.0 Operating costs $M 462.6 Cash operating cost $/t milled 86.45 Cash operating cost $/kg U3O8 produced 52.06 Total uranium production thousand kg U3O8 8,886 Total project cash flow $M 830.1 Income tax payable $M 187.4 Royalty payable $M 82.21 Project cash flow (after tax) $M 642.7 NPV @ 10%/y discount rate $M 241.5 Internal rate of return % 35 1 Deducted from revenue.

Among the factors assumed in the development of the cash flow model are:

• Mining recovery 93%, dilution 5% at zero grade. • Production rate of 1,500 t/d, operating 365 days per year. • Plant recovery 94.9%. • Sunk costs of $81.4 M have been included in the capital write off. • A royalty of 5% of revenue has been assumed. • A tax rate of 25% of taxable income has been assumed. • Analyses are in constant US dollars of mid-2007 value.

Based on these assumptions and analyses, the project cash flow yields an IRR of 35% after tax. The NPV at a discount rate of 10%/y is $241 M.

Table 1.3 shows the sensitivity of the project cash flow to uranium price, mined grade, operating costs and capital costs. As may be expected, factors which directly affect revenue, i.e., the price of uranium and mined grade, are the most sensitive items, followed in turn by capital and then operating costs. The results on IRR and net present value of varying the grade of uranium by 10% and 25% are the same as for the price of uranium.

Table 1.3 Results of Sensitivity Analysis

Factor Minus 10% Base Case Plus 10% Internal Rate Net Present Internal Rate Net Present Internal Rate Net Present of Return Value of Return Value of Return Value (%) ($ M) (%) ($ M) (%) ($ M) Uranium price 30 186 35 241 39 297 Grade of uranium 30 186 35 241 39 297 Capital cost 38 257 35 241 32 226 Operating cost 36 258 35 241 33 225 Minus 25% Base Case Plus 25% Internal Rate Net Present Internal Rate Net Present Internal Rate Net Present of Return Value of Return Value of Return Value (%) ($ M) (%) ($ M) (%) ($ M) Uranium price 22 103 35 241 45 380 Grade of uranium 22 103 35 241 45 380 Capital cost 45 282 35 241 27 201 Operating cost 38 283 35 241 31 200

The result of reducing either the price of uranium or the mined grade by 25% is an IRR of 22% which supports the robustness of the project at this level of preliminary economic assessment. The estimated grade will be further evaluated through the planned exploration program.

The results of full feasibility study of the Gurvanbulag project are likely to differ from the results of this preliminary economic assessment.

1.10 CONCLUSIONS AND RECOMMENDATIONS

It is recommended that Western Prospector/Emeelt Mines proceeds with work on feasibility studies for the Gurvanbulag project.

It is recommended, also, that:

• Exploration continues at Gurvanbulag and in the surrounding area with the objective of increasing the identified mineral resources and, potentially, increasing the life of a mining and processing operation at Gurvanbulag.

• The results of exploration through the second half of 2006 and 2007 are compiled into an updated estimate of mineral resources.

• Work is continued on development of the main access road and the power line and distribution facilities are completed.

• Work is continued on environmental baseline studies and community development programs.

• Work is continued towards securing the permits that will be required for construction and development.

• The assessment of the uranium market is updated to take account of recent movements in the spot price for uranium and its potential impact on long term contract prices that will be applicable to the Gurvanbulag project.

2.0 INTRODUCTION AND TERMS OF REFERENCE

Micon International Limited (Micon) has been retained by Western Prospector Group Ltd. (Western Prospector) to prepare a preliminary economic assessment of the Gurvanbulag property which is located in northeastern Mongolia. The following report presents the results of the preliminary economic assessment in accordance with the requirements of Canadian National Instrument (NI) 43-101.

The Gurvanbulag uranium deposit lies within the Saddle Hills property, a group of exploration licenses held by Emeelt Mines (Emeelt Mines) LLC and Western Prospector Mongolia LLC. Both Emeelt Mines and Western Prospector Mongolia are wholly-owned by Western Prospector Group Ltd., a Canadian company based in Vancouver, British Columbia. The property is operated by Emeelt Mines. The Gurvanbulag deposit was explored and partially developed in the 1970s and 1980s by the uranium exploration arm of the Ministry of Geology of the former Soviet Union.

The objective of the preliminary economic assessment was to provide an evaluation of an underground mining operation based principally on the following:

• An estimate of mineral resources for the Central Zone of the Gurvanbulag deposit prepared by SRK Consulting (Canada) Inc. (SRK), dated November 17, 2006.

• The results of metallurgical testwork carried out under the direction of Melis Engineering Ltd. (Melis) and completed in December, 2006, and a preliminary flowsheet and processing cost estimates prepared by Melis.

• The results of the mine dewatering program completed in 2006 that revealed the extent of the previously developed underground workings.

• Preliminary mining plans developed by Malcolm Buck, P.Eng., and Eugene Puritch, P.Eng.

The preliminary economic assessment report prepared by Micon will be provided to Mongolian agencies in support of application for conversion of certain exploration licenses to mining licenses.

The preliminary economic assessment is based on indicated and inferred mineral resources estimated by SRK for the Central Zone of the Gurvanbulag uranium deposit, on the results of the metallurgical testwork completed in December, 2006 and on the results of the dewatering of the underground workings at the end of 2006.

categorized as mineral reserves; there is no certainty that the preliminary assessment will be realized.

Site visits to the Gurvanbulag property have been carried out by the following individuals who have provided specific input to the present report:

Malcolm Buck January 29-30, 2005. Bruce Fielder, Melis October 13-25, 2006, June 15-20, 2007. Chris Lee, SRK February 1-4, 2006. Eugene Puritch, P&E November 14-18, 2006.

The Qualified Persons who have prepared this report are:

Malcolm Buck, P.Eng., Associate Mining Engineer, Micon International Limited Bruce C. Fielder, P.Eng., Senior Processing Engineer, Melis Engineering Ltd. Marek Nowak, MASc., P.Eng., Principal Geostatistician, SRK Consulting (Canada) Inc. Eugene Puritch, P.Eng., President, P&E Mining Consultants Inc. Jane Spooner, P.Geo., Principal, Micon International Limited Mani M. Verma, P.Eng., Associate Mining Engineer, Micon International Limited

Each of the Qualified Persons listed above is independent of Western Prospector/Emeelt Mines as set out in Section 1.4 of NI 43-101.

2.1 ACKNOWLEDGEMENTS

Micon acknowledges and appreciates the input of the individuals and firms noted above. In particular, Micon appreciates the assistance that was provided by Dr. Gerald Harper and Mr. Bruce Brady of Western Prospector.

2.2 UNITS OF MEASURE AND ABBREVIATIONS

Metric units of measure have been used throughout this report unless noted otherwise. For example, prices for uranium may be expressed in terms of United States dollars per pound of contained uranium oxide (U3O8). Costs have been estimated in United States dollars ($) of mid-2007 value.

A list of abbreviations is provided in Table 2.1.

Table 2.1 List of Abbreviations

EcoTrade LLC EcoTrade Emeelt Mines LLC Emeelt Mines Golder Associates Golder Melis Engineering Ltd. Melis Micon International Limited Micon

Mineral Resources Authority of Mongolia MRAM P&E Mining Consultants Inc. P&E SRK Consulting (Canada) Inc. SRK Western Prospector Group Ltd. Western Prospector Acceleration due to gravity g Ampere(s) A Bequerel Bq Bequerel per litre Bq/L Billion years ago Ga Bond Work Index BWI Canadian dollar(s) Cdn$ Centimetre(s) cm Copper Cu Cost and freight cfr Cubic feet per minute cfm Cubic metre(s) m3 Cubic metre(s) per day m3/d Cubic metres per second m3/s Cubic yard(s) yd3 Degree(s) o Degree Celsius oC Effective grinding length EGL Foot (feet) ft Gallons per minute gpm Gigawatt(s) GW Gigawatts electrical GWe Gigawatt hour(s) GWh Grams g Grams per tonne g/t Grauss Kruger GK Hertz Hz High density polyethylene HDPE Horsepower HP Horsepower per tonne HP/t Hour(s) h Hour(s) per day h/d Inch(es) in In-situ leach ISL Kilometre(s) km Kilometres per hour km/h Kilopascal(s) kPa Kilovolt kV Kilovolt ampere(s) kVa Kilowatt(s) kW Kilowatt hours per tonne kWh/t Litre(s) L Litres per second L/s Megawatt(s) MW Megawatts electrical MWe Metre(s) m Metres above sea level masl Metres per second m/s

Micron(s) μ, mμ Milligram(s) mg Milligrams per litre mg/L Millilitre(s) mL Millimetre(s) mm Millimetres per year mm/y Million M Million pounds Mlb Million cubic metre(s) Mm3 Million litres ML Million litres per year ML/y Million tonnes Mt Million years ago Ma Minute(s) Min National Instrument 43-101 NI 43-101 Net smelter return NSR Non-governmental organization(s) NGO(s) Oxidation reduction potential ORP Parts per million ppm Parts per billion ppb Per cent % Percent by weight w/w Pound(s) lb Quality assurance/quality control QA/QC Rock quality designation RQD Second s Semi-autogenous grinding SAG Specific gravity SG Square metre(s) m2 Square kilometre(s) km2 Standard penetration test SPT Tonne(s) t Tonnes per day t/d Tonnes per hour t/h Tonnes per year t/y Total suspended solids TSS United States dollars $ Universal Transverse Mercator UTM Volt(s) V X-ray diffraction XRF Year(s) y

3.0 RELIANCE ON OTHER EXPERTS

In preparing the preliminary economic assessment of the Gurvanbulag property, Micon has relied upon the following:

• The Report on Saddle Hills area uranium prospects, Exploration Licenses 3367X, 4846X, 4969X, 4974X, 6995X, 7023X, 7405X, 7685X and 9283X, Dornog Aimag, eastern Mongolia, prepared by Gerald Harper on behalf of Western Prospector, dated May 4, 2005. This report was filed on SEDAR on May 11, 2005.

• The Independent Technical Report and Resource Estimate for the Gurvanbulag Deposit, Saddle Hills Uranium Project, Mongolia prepared by SRK on behalf of Western Prospector, dated November 17, 2006. This report was filed on SEDAR in June, 2007.

• The results of metallurgical testwork carried out under the direction of Melis at SGS Lakefield Research Limited (SGS Lakefield) as described in a memorandum from Melis to Western Prospector dated February 14, 20007, and a preliminary flowsheet developed by Melis and capital and operating cost estimates prepared by Melis. Melis was retained for this work by Western Prospector.

• Preliminary mining plans developed by Malcolm Buck, P.Eng., with input from Eugene Puritch, P.Eng. Mr. Buck was retained for this work by Micon. Mr. Puritch of P&E Mining Consultants Inc. (P&E) was retained by Western Prospector.

• An assessment of the market for uranium prepared by Colorado Nuclear, Inc. (CNI). CNI was retained by Western Prospector

• Data and certain estimates of cost provided for manpower and infrastructure elements of the project by Western Prospector.

Micon has not carried out any exploration work, drilled any holes or carried out any sampling or assaying of material from the Gurvanbulag property.

While exercising all reasonable diligence in checking and confirming it, in the preparation of this report Micon has relied upon the data provided by individuals and companies, as noted above, and by Western Prospector/Emeelt Mines, and that found in public domain sources.

The status of the licenses under which Emeelt Mines, Western Prospector Monglia and Adamas Mining LLC (Adamas) hold title to the surface and mineral rights for the Gurvanbulag property has not been investigated by Micon, and Micon offers no opinion as to the validity of the title claimed by Emeelt Mines, Western Prospector and Adamas. The description of the property, and ownership thereof, as set out in this report, is provided for general information purposes only.

4.0 PROPERTY DESCRIPTION AND LOCATION

The Saddle Hills area is located in Dornod Aimag in eastern Mongolia, approximately 100 km from the border of Mongolia with Russia to the north, and a similar distance from the border with China to the east. See Figure 4.1.

Eleven exploration licenses are held directly by Western Prospector Mongolia LLC and Emeelt Mines LLC, as shown in Table 4.1. Both Emeelt Mines LLC and Western Prospector Mongolia LLC are wholly-owned by Western Fortune Ltd., a Malaysian-registered company wholly-owned by Western Prospector Group Ltd. One license is held by Adamas Mining LLC (Adamas) in which Western Prospector Mongolia LLC is earning a 70% joint venture interest.

Figure 4.1 General Location Map

Harper (2005) reported:

“WNP [Western Prospector] contracted a land survey in September, 2004 and subsequently made application for various areas of surface rights to cover lands

contemplated as needed for surface facilities related to future mining operations, infrastructure and transportation corridors.”

Harper (2005) describes the system of license ownership and provides a tabulation of the geographic specifications of the Saddle Hills area exploration licenses to which the reader is referred.

Table 4.1 Saddle Hills Area Exploration Licenses

License Area name Aimag Soum Surface Owner Interest Number (ha) (%) 3367X Erkht ovoot tolgoi Dornod Dashbalbar, 39,942 Adamas Mining Earning Gurvanzagal LLC 70% JV interest 4846X Burd Dornod Dashbalbar, 5,753 Emeelt Mines 100% Bayandun 4969X Ulaan Nuur 2 Dornod Sergelen, 6,761 Emeelt Mines 100% Bayandun, Dashbalbar 4970X Ulaan Nuur 1 Dornod Sergelen, 10,876 Emeelt Mines 100% Bayandun 4974X Ulaan Nuur 3 Dornod Dashbalbar, 15,092 Emeelt Mines 100% Sergelen 6995X Bulagtai Dornod Bayandun 1,017 Emeelt Mines 100% 7023X Baishintiin ukhaa Dornod Dashbalbar, 12,963 Emeelt Mines 100% Gurvanzagal 7405X Dalan khariin bulag Dornod Sergelen 1,031 Emeelt Mines 100% 7685X Ar bulag Dornod Bayandun 7,254 Western 100% Prospector Mongolia 9283X Dashbalbar 2 Dornod Dashbalbar 127 Emeelt Mines 100% 9353X Khureet nuur Dornod Tsagaan-Ovoo, 52,811 Western 100% Bayandun, Prospector Sergelen Mongolia 9363X Ugtam Uul Dornod Dashbalbar 34,939 Western 100% Prospector Mongolia

License 7685X, held by Western Prospector Mongolia, covers a significant part of the area planned by Western Prospector/Emeelt Mines for potential initial uranium production. Emeelt Mines owns license 4969X which covers the balance of the area in which mineral resource blocks have been defined relating to the Gurvanbulag uranium deposit. Licenses 6995X and 7405X cover the area down dip from the open-ended limit of drill defined resources. Emeelt Mines plans to apply for mining licenses over parts of the following licenses in respect of the Gurvanbulag deposit:

7685X 4969X

6995X 7405X

Emeelt Mines is also exploring several additional uranium deposits within the property area and, on completion of mineral resource estimations, may apply for mining licenses over other exploration licenses including 3367X, 4974X, 9283X and 9353X.

The licenses are contiguous, as shown in Figure 4.2, and cover a total area of 188,566 ha. The extent of the area covered by the licenses is approximately 83 km in a north-south direction and 80 km in an east-west direction. The centre of the licensed area is at latitude 49o N and longitude 114o E. The Main shaft is located on license 7685 approximately at latitude 49o 05” N and longitude 114o E.

Also shown on Figure 4.2 are the State Protected Areas of Ugtam Uul in the northwest and Yahi Nuur in the south.

Within the area of the Western Prospector and Emeelt Mines exploration licenses, are several small licenses owned by third parties. At about 49o05’ N and 114o05’ E, licenses 247A and 1562X are owned by the Chinese mining company, XinXin Mining Co. Ltd. (XinXin), which is developing the Ulaan polymetallic zinc, lead, silver deposit. Further east at approximately 49o05’ N and 114o20’ E, mining license 237A covers the Dornod uranium deposits and is owned by Central Asia Uranium Corporation which is 58% owned by Khan Resources Inc., a Canadian company. Immediately to the south of 237A, Khan Resources Inc. owns 100% of exploration license 9282X subject to a 3% royalty to Western Prospector.

Exploration license 3367X is subject to an option and joint venture agreement between the initial owner, Adamas, and Western Prospector which, having fulfilled its work expenditure requirements, is in process of converting its earned interest to a 70% ownership in a joint venture company to be formed. Adamas will retain a 30% joint venture interest at the outset and will also receive a 3% royalty on any future production from the license. It covers an area of 39,942 ha and is centred at approximately 49o10’ N and 114o15’ E. See Figure 3.2.

Western Prospector has advised Micon that a royalty of 5% of revenue is payable to the Government of Mongolia.

Table 4.2 shows the ownership, map locations and issue dates for the exploration licenses. Emeelt Mines and Western Prospector have 100% ownership in the licenses, except as noted, in Table 4.2.

Figure 4.2 Location of Exploration Licenses

On September 11, 2007 Western Prospector and Emeelt Mines were notified by the Mongolian Mineral Resources Administration that because the Gurvanbulag and other deposits had been discovered with State Funds as defined in the Mongolian Mineral Law (1996) and had previously defined reserves, they should be converted to mining licenses as a priority. Western Prospector has initiated the application for conversion.

Table 4.2 Details of Exploration Licenses

Exploration Map Sheet Location Issue Date Area Ownership Next License License (ha) Payment ($) Number 3367X M-50-97 07-Jun-01 39,942 Adamas Mining LLC1 59,913.00 4846X M-49-108 05-Sept-02 5,753 Emeelt Mines LLC 5,753.00 4969X M-50-97 & M-50-109 08-Oct-02 6,761 Emeelt Mines LLC 6,761.00 4970X M-49-120 08-Oct-02 10,876 Emeelt Mines LLC 10,876.00 4974X M-50-97 & M-50-85 08-Oct-02 15,092 Emeelt Mines LLC 15,092.00 6995X M-50-97 25-Feb-04 1,017 Emeelt Mines LLC 1,017.00 7023X M-50-98 04-Mar-04 12,963 Emeelt Mines LLC 12,963.00 7405X M-50-97 07-May-04 1,031 Emeelt Mines LLC 1,031.00 7685X M-50-96 10-Jun-04 7,254 Western Prospector Mongolia LLC 7,254.00 9283X M-50-97 11-Feb-05 127 Emeelt Mines LLC 38.10 9353X M-50-109 & M-49-120 23-Feb-05 52,811 Western Prospector Mongolia LLC 15,843.30 9363X M-49-96 & M-49-108 07-Mar-05 34,939 Western Prospector Mongolia LLC 10,481.70 1 Western Prospector Mongolia is earning a 70% joint venture interest.

The licenses shown in Figure 4.2 cover part of five soums, Dashbalbar, Gurvanzagal, Sergelen, Bayandun and Tsagaan-Ovoo. Within the aimags, or primary administrative areas of Mongolia, soums are the secondary administrative areas.

The Gurvanbulag deposit is located on the border between Sergelen and Bayandun soums and Western Prospector/Emeelt Mines have acquired surface rights to cover lands that may be needed for surface facilities related to future mining operations, on-site processing facilities, infrastructure facilities and transportation corridors. The surface rights are authorized by the aimag government following approval by the responsible soum government. Surface leases are maintained by payment of annual rents at escalating rates per hectare.

The Mineral Resources Authority of Mongolia (MRAM), the agency responsible for mineral licensing, uses the UTM system of coordinates. Topographic maps of Russian origin use the Gauss Kruger (GK) coordinate system which has no direct correlation with the UTM system. Western Prospector uses the GK system.

4.1 ENVIRONMENTAL CONSIDERATIONS

Work undertaken by Western Prospector/Emeelt Mines relating to environmental baseline studies for the Gurvanbulag property and for construction of infrastructure facilities are described in Section 18.10 of this report. Micon is not aware of any environmental liabilities associated with the property that are not recognized and being managed by Western Prospector/Emeelt mines through its programs as described in Section 18.10.

4.2 PERMITS

As far as Micon is aware, Western Prospector/Emeelt Mines have obtained all necessary permits to conduct its exploration and dewatering programs on the Gurvanbulag property, and for development of infrastructure facilities. See also Section 18.10 of this report.

5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

A railway line runs north from Choibalsan via Boorj in Russia to connect with the Trans- Siberian Railway line. Rail track remains on the branch line to the Dornod area but is unused, and a graded rail bed runs from there through Mardai and on to Gurvanbulag.

The Saddle Hills area receives an average of 250 mm/y precipitation as rain and snow. Temperatures in Dornod Aimag average -20.5oC in January and 19.9oC in July. The area is characterized by high rounded hills, up to 1,000 masl, and generally flat alluvial plains. Rivers cut steeply sided valleys at the higher elevations. Surface water remains year-round in the larger rivers and lakes. The water table is relatively close to surface in the valleys and is accessible by herders using wells some 2 m deep. At higher elevations, however, the water table may be over 100 m below surface.

Average precipitation at the Bayandun weather station is shown in Table 5.1.

Table 5.1 Average Annual Precipitation at Bayandun Weather Station (Millimetres)

1994 1995 1996 1997 1998 1999 2000 2001 2002 2003 2004 301.3 286.0 416.7 115.2 642.1 242.2 280.1 160.9 380.5 230.6 146.8

Western Prospector/Emeelt Mines have operated a weather station at the Gurvanbulag site and have collected ground and air temperatures, wind speed and precipitation data since May, 2005.

Although winter temperatures are relatively low and winds may be strong, Western Prospector/Emeelt Mines conduct activities at site throughout the year.

Wildlife includes large herds of antelope, small packs of wolves and a variety of small burrowing mammals. Numerous bird species are present year-round and the area is visited by migrating birds.

The local population comprises nomadic herders who raise horses, cattle, sheep, goats and camels. Their activities are mostly restricted to the flat valley floor areas where water is available at or near surface.

The town of Mardai was constructed during the period of Russian exploration and development of the Saddle Hills uranium district. Mardai was abandoned in the early-1990s and infrastructure, including the railway line, has been dismantled and sold for scrap.

The largest regional centre of population is the town of Choibalsan which has a population of approximately 39,000. Several higher education facilities are located in Choibalsan, including the Mongolian Knowledge University. One so-called general education school is located in each of the soum centres of Gurvanzagal, Sergelen, Dashdalbar and Bayandum. The nearest hospital to Gurvanbulag is located in Choibalsan.

XinXin is developing the Ulaan polymetallic deposit located 7 km east of Gurvanbulag in Dashbalbar soum. XinXin has purchased from Emeelt Mines a small area of license 4974X to provide it with space for mill and tailings facilities. XinXin and Emeelt Mines have a 50:50 joint venture agreement to build and own a high voltage powerline to deliver power from the Choibalsan thermal power station to their project sites. This line and its substations are under construction.

In 2007, a joint venture agreement was concluded between Emeelt Mines and XinXin to provide for construction and operation of a powerline to their respective project sites. Engineering, design, environmental assessment and permitting were completed and construction is currently underway. The estimated completion date is October, 2007.

Engineering for an upgraded road from Choibalsan to the mine site was completed and filed with the Ministry of Roads. Environmental assessment studies have been initiated leading to permit applications for road construction.

Emeelt Mines is providing the government of Mongolia with information related to the project preparatory to initiating discussions for an Economic Benefits Agreement. Formal discussions for the Economic Benefits Agreement will not start until approval of the State Reserves Committee has been received and application made for the exploration licenses to be converted to mining licenses.

6.0 HISTORY

The Gurvanbulag property, located in the Saddle Hills area of northeastern Mongolia, was explored and developed by the Ministry of Geology of the Soviet Union in the 1970s and 1980s. It was abandoned in the early 1990s following the collapse of the former Soviet Union. The shafts were capped, surface infrastructure was removed or destroyed and the partially developed underground workings were allowed to flood.

Harper (2005) provides a synthesis of the historical work undertaken by the Russians. This is also described in SRK (2006).

Exploration work was conducted by the Russians between 1944 and 1989. This work, and the historical uranium resource estimates that were prepared using the Soviet system of mineral resource and reserve classification, are described by Harper (2005). The historical resource estimates are not in compliance with NI 43-101.

Regional geological exploration was undertaken by the Russians from 1944. Polymetallic mineralization was identified in 1945. Uranium anomalies were identified in the early-1970s and work continued with drilling and, in the early-1980s, the Gurvanbulag deposit was developed by means of excavation of three shafts and extensive lateral underground development. Following the collapse of the former Soviet Union, the Gurvanbulag mine was allowed to flood, all surface facilities relating to the development of the Gurvanbulag deposit were removed and all shafts were capped with concrete.

Western Prospector/Emeelt Mines started to acquire exploration licenses in the Saddle Hills area in 2004 and initiated an exploration program that was focused on the Gurvanbulag deposit. Harper’s review of the available Russian data led to the decision by Western Prospector/Emeelt Mines to dewater the Gurvanbulag mine in the second half of 2006 and to initiate underground exploration and sampling. Western Prospector/Emeelt Mines believe that the Russian drill core and samples no longer exist.

7.0 GEOLOGICAL SETTING

The geology and mineral resources for the Gurvanbulag property are described in two principal documents:

• Report on Saddle Hills area uranium prospects, Exploration Licenses 3367X, 4846X, 4969X, 4970X, 4974X, 6995X, 7023X, 7405X, 7685X and 9283X, Dornod Aimag, eastern Mongolia, for Western Prospector Group Ltd., May 4, 2005 by Harper, G. (2005).

• Independent Technical Report and Resource Estimate for the Gurvanbulag Deposit; Saddle Hills Uranium Project, Mongolia, November 17, 2006 by SRK Consulting (Canada) Inc. (SRK (2006)).

Both reports were prepared in compliance with Canadian National Instrument (NI) 43-101. Harper’s report was filed on SEDAR in May, 2005. The SRK report was filed on SEDAR in June, 2007 (www.sedar.com).

The following descriptions have been summarized from SRK (2006) and Figures 7.1 and 7.2 are reproduced from the SRK report.

7.1 REGIONAL GEOLOGY

Mongolia occupies part of the Central Asian Fold Belt. In the literature, the belt is also variously called the Central Asian Mobile Belt, Central Asian Mountain Belt, Central Asian Orogenic Belt and Ural-Mongolian Fold Belt. The Central Asian Fold Belt is juxtaposed against the Siberian craton to the north, and the Sino-Korean craton to the south.

Rocks as old as 2.65 Ga have been found. Repeated cycles of sedimentation, metamorphism, rare mafic intrusions, abundant granitic intrusions, felsic extrusions, orogenic cycles and plate tectonic events mark an ongoing geological history with few apparent long breaks. Uranium values enriched to several times crustal abundance are common in the youngest felsic volcanic rocks of Mesozoic age and these are believed to have provided the sources of uranium mineralization at Strel’tsov in Russia and Saddle Hills in northeastern Mongolia.

7.1.1 Central Mongolia Fold Belt

Based on the work of many Mongolian and Russian geologists, Mironov divides Mongolia into seven fold belts or systems (see Figure 7.1):

• Mongolia Altai. • North Mongolia. • Transbaikal-Mongolia. • Central Mongolia.

• South Mongolia. • South Gobi. • Inner Mongolia.

Figure 7.1 Geological Fold Belts in Mongolia

The Saddle Hills area is part of the Central Mongolia fold belt. The belt is characterized by Proterozoic gneisses, schists, and anatectic granites, occurring as inliers in early Paleozoic volcanic and sedimentary rocks. Silurian and Devonian sandstones and shales are widespread. Upper Paleozoic volcanic and molasse-type sedimentary rocks unconformably overlie earlier Proterozoic and Paleozoic rocks.

7.1.2 Argun-Mongolia province

Mironov further divides Mongolia into a hierarchy of uranium ore ‘provinces’, ‘districts’ and ‘clusters’. There are four uranium ore ‘provinces’, named the Argun-Mongolia, Gobi- Tamisag, Kentii-Daur, and North Mongolia provinces (Figure 7.2). The Saddle Hills area lies within the Argun-Mongolia province which contains four uranium ‘districts’. From southwest to northeast these districts are named the Central Gobi, East Gobi, Berh, and North Choibalsan, which includes the ‘Dornod cluster’. The Saddle Hills area with its Gurvanbulag, Dornod and Mardaingol deposits and occurrences forms the ‘Dornod uranium ore cluster’ of the Argun-Mongolia province.

Figure 7.2 Uranium Provinces and Districts of Mongolia

7.1.3 Stratigraphic Setting

The Argun-Mongolia uranium province is characterized by bimodal continental basalt- rhyolite rocks and interbedded continental sedimentary rocks throughout its length. The three uranium mineralized sites in the Saddle Hills area, i.e., Gurvanbulag, Dornod and Mardaingol, occur in volcanic rocks, filling what appear to be volcanic depressions, or down- faulted crustal blocks. The stratigraphic profiles at the three sites are similar but not identical. Each has a lower succession of sedimentary rocks, basalts, intermediate tuffs and rhyolites, capped by a thick upper succession of felsic volcanic rocks.

The volcanic rocks erupted during the late Jurassic to early Cretaceous Periods. Imprecise potassium-argon dates indicate that the time in which the volcanic rocks erupted may have spanned 5-35 million years.

The geological structure of the Gurvanbulag deposit comprises two structural stages, the basement rocks of the lower stage and Mesozoic rocks of the upper stage. Uranium mineralization has been located in the upper structural stage but, by analogy with Strel’tsov, may occur in the lower structural stage as well.

The reader is referred to Harper (2005) and SRK (2006) for more detailed descriptions of the geological setting of the Gurvanbulag deposit.

8.0 DEPOSIT TYPES

The following summary of the types of uranium deposit is based on SRK (2006) to which the reader is referred for a more detailed description of the classification of uranium deposits.

The majority of uranium production, world wide, has come from three principal types of deposit: pyritic quartz pebble conglomerates, sandstone, and unconformity. With the closing of the Elliot Lake mines in Canada and the downscaling of gold mining in the Witwatersrand of South Africa, where uranium is a by-product, production is now primarily from the unconformity and sandstone types.

8.1 URANIUM IN VOLCANIC ROCKS

Uranium deposits in volcanic rocks make up only a small proportion of the world’s uranium resources. They appear to be more prominent and more common in territories of the former Soviet Union where, for Russian geologists, the association of uranium-fluorite-molybdenum is a dominant type. SRK (2006) provides a listing of some of the better known volcanic- hosted uranium occurrences as follows:

• Michelin deposit in the Central Mineral Belt of Labrador, Canada. • McDermitt caldera in northwestern Nevada, United States. • Margarita Mine in Mexico. • Ben Lomond and Maureen in Queensland, Australia. • Tianmujian caldera, Jaingxi province, China. • Occurrences in Kazakhstan. • Strel’tsov complex, Russia • Dornod complex, Mongolia.

Uranium deposits in volcanic rocks occur as hydrothermal veins in steeply dipping faults and shear zones which may cut a complex volcanic-sedimentary stratigraphy or its immediate basement. Bedding conformable deposits, if present, are generally small. Complex and intense hydrothermal alteration is always present.

The Saddle Hills district is distinctive in the presence of the laterally extensive volcanic obsidian horizon below rhyolites at Gurvanbulag and the laterally extensive uranium mineralization that is conformable to bedding in parts of this horizon. SRK (2006) notes that, in this respect, the Gurvanbulag uranium deposit is not comparable to any other economic or potentially economic uranium deposit.

9.0 MINERALIZATION

As noted in Section 7.1.3, above, the geological structure of the Gurvanbulag deposit comprises two structural stages, the basement rocks of the lower stage and Mesozoic rocks of the upper stage. Uranium mineralization has been located in the upper structural stage but, by analogy with Strel’tsov, may occur in the lower structural stage as well.

More detailed descriptions of the mineralization at Gurvanbulag are provided in Harper (2005) and SRK (2006).

9.1 BASEMENT ROCKS (LOWER STRUCTURAL STAGE)

The Proterozoic basement rocks include amphibolites, biotite-hornblende and hornblende crystalline slates and schists, biotite-muscovite and sillimanite gneisses with intercalations of quartzites, calciphyres (calcsilicate-marble) and marbleized limestones. Among the Paleozoic formations are gabbro-diorites, diorites, granodiorites, and leucocratic and biotite granites with many conformable contacts with the host metamorphic rocks and characterized by retaining the original gneiss foliation and by the presence of xenoliths of unaltered substrate rocks.

Borehole and geophysical data show that the basement relief is complex, due to erosion and tectonic processes. Given its generally gentle dip to the southeast at angles of 10-30o, various basement depressions and elevations have been identified, differing in scale, form and orientation. The largest of the depressions can be traced in the basement across the eastern part of the Gurvanbulag deposit in the form of a broad (1.0-2.5 km) flat-bottomed paleovalley. Towards the north it is bounded by a projection of the basement, and in the south it merges with the Tukhemiynskaya Basin. It branches into a series of smaller trough- like depressions between which are dome- and ridge-shaped basement elevations. The Gurvanbulag Central and Southwest zones are located over the slopes of higher paleo- elevations.

9.2 MESOZOIC ROCKS (UPPER STRUCTURAL STAGE)

The Upper Mesozoic (late Jurassic to early Cretaceous) sedimentary-volcanic deposits of the Dornod complex are divided into the lower, middle and upper subcomplexes, separated by breaks and unconformities.

The lower subcomplex is an intercalation of layers of quartz-feldspar porphyries, trachydacites and andesite-basalts with horizons of tuff-sedimentary rocks. The total thickness of the subcomplex varies from 150 to 450 m.

The middle subcomplex comprises structurally and texturally varied layers of volcanic rocks, with a total thickness of 300-800 m.

The upper subcomplex comprises dominantly felsites or rhyolites but includes layers of trachyandesites a few hundred metres thick. This sedimentary-volcanic sequence has a

gentle monoclinal dip to the southeast of 5-20°. Near the margin and around large faults, the dip increases to 30-40°. The stratigraphic column of the Gurvanbulag deposit is presented in Table 9.1 and has been reproduced from SRK (2006).

Table 9.1 Stratigraphic Succession in the Saddle Hills Area

Maximum Unit Description Thickness (m) 200 Upper Subcomplex microlitic trachyandesites 10 conglomerates, sandstones ~~~~~ possible disconformity ~~~~~ 600 sanidine small prophyritic trachyrhyolites, their lava breccias 15 volcanic glass deposits (obsidians) (level of the Khurtynbulag fault) 250 flow and spherulitic felsites 50 vitroclastic, ashy tuffs, obsidians (level of the Khayakh fault) 700 massive coarse-flow felsites, at the bottom – spherotaxitic, at the top – spherulitized with lendes of tuffs, tuff-breccias, and volcanic Middle Subcomplex glass deposits (obsidians. 15 volcanic glass deposits (obsdidians) (level of the Gurvanbulag deposit) 150 vitroclastic ashy felsite tuffs 200 oligophyric rhyolites, their lava breccias and tuffs 150 vitrolithic felsite tuffs 30 conglomerate, sandstones, tuff-sandstones, siltstones ~~~~~ possible disconformity ~~~~~ 30 amydaloidal andesite-basalts (3rd layer) 15 conglomerates, sandstones, siltstones 160 trachydacites, their lava breccias and tuffs 20 sandstones, tuff-standstones 120 Lower Subcomplex massive amygdaloidal andesite-basalts (2nd layer) 15 sandstones, tuff-sandstones 140 quartz-feldspar porphyries, their tuffs 70 dacite tuffs, tuffites 130 conglomerates, sandstones, siltstones, with lenses of acid tuffs ~~~~~ major unconformity ~~~~~ Proterozoic crystalline slates, schists and gneisses, amphibolites, diopside[?]-scapolite quartzites, calciphyres, marbleized limestones Basement with Paleozoic intrusion-anatectoid formations of granite- granodiorite and diorite-granite formations

The Gurvanbulag deposit occurs in close spatial association with an obsidian (volcanic glass) horizon within the middle sub-complex. This is reported as being up to 15 m thick in the Russian literature although drill hole data indicate a maximum of 12 m and, generally, a thickness of 3-8 m. Mineralization is primarily hosted in the vitroclastic, ashy felsite tuffs in the immediate footwall to the obsidian horizon, but is also locally well-developed in the hanging wall felsites. The obsidian layer, itself, is variably mineralized and is commonly barren.

The ashy felsite tuffs range from well-bedded ash, to fine lapilli tuffs, to coarse fragment agglomerates, to a mottled, purple and beige indefinable rock without clear clasts. The latter rock appears to have fused under high temperatures.

The obsidian occurs at the base of the upper felsites. Thin beds of tuffs may intervene, but mostly the obsidian is in contact with felsites. The obsidian erupted about 150 Ma ago. During this time it has mostly devitrified, but many hand specimens look quite fresh.

The least altered obsidian is vitreous black with a conchoidal fracture. All stages of devitrification to complete replacement by light grey to white, putty like clay can be observed. The obsidian horizon is a laterally extensive, relatively thin sheet which can be traced in outcrop for 15-20 km and extends an unknown distance down dip. Mironov reports an area of 159 km2 for the Gurvanbulag level.

9.3 STRUCTURE

The Gurvanbulag, Dornod, Mardaingol uranium deposits occur in three tectonic “blocks”, respectively known as Ulaan, Erkh(tiyskiy) and Mardaingol. The blocks are separated from each other by prominent faults, some with measurable offsets of stratigraphic contacts. Dornod lies about 25 km east of Gurvanbulag while Mardaingol is about 6 km northwest of Dornod. Geologically, Dornod and Mardaingol, being close together, have more in common with each other than either has with the more distant Gurvanbulag deposit.

Intrablock faulting in the Ulaan block, and specifically in the Gurvanbulag deposit, is prominent. Faults are of two kinds, steeply-dipping and shallow-dipping. Steeply-dipping faults are abundant and penetrate the entire stratigraphic succession. Laterally, they may be traced for great distances.

Shallow-dipping faults are conformable to bedding. The Gurvanbulag mineralization locally coincides with, and is interpreted by the Russians as having been controlled by, such faulting.

9.4 HYDROTHERMAL ALTERATION

Most studies of alteration in the Saddle Hills area were carried out by the Russians in the 1970s and 1980s, as noted by Harper (2005). Broadly, alteration has been divided into two categories:

• High-temperature alteration in basement rocks that predates most of the volcanic cover. (Only at Ulaan is high temperature alteration found within volcanic rocks).

• Low-temperature alteration and introduction of uranium after late Mesozoic/early Cretaceous volcanism had ceased.

Late Jurassic to Early Cretaceous alteration resulted in a suite of minerals including hydromicas, mixed layer micas, montmorillonite, carbonates, kaolinite, chlorite, feldspars,

sericite, epidote, uranium minerals (coffinite, pitchblende and uranophane), fluorite and molybdenite.

Russian reports describe widespread, pervasive hydromica alteration that may affect more than half of the volcanic rocks and intense localized alteration in the uranium mineralized zones. Formation of hydromicas further increased porosity of the already preferred porous pathways, thus preparing the ground for the following stage of uranium-bearing hydrothermal solutions. The most strongly developed hydromica alteration does not necessarily contain uranium, but uranium always occurs with strong hydromica development.

The main uranium minerals are, in decreasing order of abundance, coffinite, uranophane and uraninite/pitchblende. Alteration in the principal mineralized zone at Gurvanbulag consists of white, putty-like clay minerals, silicification, development of deep red hematite, red feldspars, chlorite, carbonates, albite, analcite and fluorite.

10.0 EXPLORATION

Exploration at Gurvanbulag has been described in Harper (2005) and SRK (2006).

10.1 2005 EXPLORATION PROGRAM

Following the acquisition of the licenses in the area and the preliminary work carried out in 2004 by Western Prospector/Emeelt Mines, the objectives for the 2005 exploration program, as defined in Harper (2005), became:

• To enhance the confidence in the knowledge of the Inferred Resources to allow them to be reviewed for definition as either Indicated or Measured Resources.

• To enhance confidence in much, or all, of the existing C2 category of Russian resources to assess its compatibility with Inferred or higher categorization of Resources.

• Test targets defined by airborne radiometric survey as having potential for deposits with immediate priority being given to those suggestive of shallow deposits which could result in definition of resources amenable to open pit mining.

The proposed exploration program included surface drilling, as well as underground dewatering and sampling of the Gurvanbulag Central deposit, in order to allow the completion of an NI 43-101 compatible resource estimation. Largely due to delays in permitting, the resource estimation and the underground dewatering and sampling of the Gurvanbulag deposit had to be postponed to 2006. The 2005 exploration program included:

• Setting up a base camp and an office at the Gurvanbulag mine site.

• Ground radiometric (1,533 line-km) and magnetic (901 line-km) surveys covering anomalous areas identified in the 2004 airborne radiometric survey.

• Detailed geological mapping of the grids (87 km2) used for the geophysics.

• Trenching, dug both by hand and excavator.

• Sampling of the Gurvanbulag mineralized dump.

• Digitization of all Russian drill holes for the Gurvanbulag deposit.

• Diamond drilling (16,488.85 m).

10.2 2006 EXPLORATION PROGRAM

The SRK (2006) technical report and resource estimate were based on drill hole data up to mid-March 2006.

In 2006, work comprised:

• Continued surface drilling in the vicinity of the Gurvanbulag deposit until September, 2006.

• Dewatering of and access to the underground mine workings being achieved in October, 2006 to the 260 m level at a depth of 260 m below surface.

• Sampling, surveying and mapping of the underground mineralized zones starting in November, 2006 and continuing through 2007. The intense clay alteration associated with some intersections through the mineralized horizon resulted in ground collapse, which has required rehabilitation. A series of several hundred percussion holes drilled from underground into the mineralized horizon for which no Russian results were available, are in process of being spectrometrically probed to provide assay values.

A total of 98 surface drill holes, totaling 31,962 m, were completed in 2006.

A limited soil sampling program was also undertaken, as was an Alpha Cup radon survey. As a result of this work, there appears not to be any easily detectable long distance halo effect.

The interpretation of exploration data is provided in Harper (2005) and SRK (2006).

Drilling and exploration conducted on the Gurvanbulag Central zone after March 18, 2006 was not utilized for the compilation of the mineral resource estimate by SRK and, therefore, was not included in the Micon preliminary economic assessment.

11.0 DRILLING

The 2005 drilling program is described in SRK (2006). Summary statistics are provided in SRK (2006) as shown in Table 11.1.

Table 11.1 Summary Statistics of 2005 Saddle Hills Drilling Program

LandDrill 38 AIDD Diamec U6 AIDD CS1000 Start date 21-Jun 11-Jul 11-Sep End date 11-Dec 10-Dec 12-Dec Number of shifts 346 308 183 Metres drilled (m) 9,260 4,784 2,435 Average rate (m/shift) 26.8 15.5 13.3 Average rate (m/d) 53.5 31.1 26.6

The following description of drilling is reproduced from SRK (2006).

“The drillers were instructed not to use any drill mud containing either molybdenum or barium, elements analyzed in the core. Molybdenum and barium are considered deleterious to the metallurgy of the ore.

“Numbering of drillholes was set up prefixed by a three letter alphabetical identifier of the area of drilling, followed by a four numbers sequence. Prefix letters used in 2005 include: • Gurvanbulag Central area = GCE • Gurvanbulag Southwest area = GSW • Mardaingol Northwest area = MNW • Mardaingol Southwest area = MSW • Mardaingol Southeast area = MSE

“The four numbers sequence used in 2005 was as follows: • Gurvanbulag area (i.e. west of a north-south line roughly through Mardai): all surface drill holes were numbered in the 5000 series (from 5001 to 5078). • Mardaingol area (Mineral License 3367X): all surface drillholes were numbered chronologically in the 7000 series (from 7001 to 7005). • Any surface drillholes in the Dornod area (i.e. within the Mineral Licenses excluding 3367X east of a north-south line roughly through Mardai) would have been numbered chronologically in the 9000 series.

“When a drill hole was for some reason abandoned and restarted from surface, the second hole was given a separate number, without any consideration on the closeness of the collars. For instance, GCE5061 and GCE5062 were re-drills of GCE5036, and GCE5063 was a re- drill of GCE5039. A wedged drillhole should have a suffix after its number. In 2005, the rule was not applied to the wedge of drillhole GSW5060, which was named GSW5069 instead of GSW5060A.

“For the Gurvanbulag area, 13,820.45 metres of drilling were completed in the Central Zone and 1,752.70 metres in the South West zone. In the Mardaingol area, 905.5m were completed

in the three zones: 411 metres in the South West Zone, 133.5 metres in the North West Zone and 361.0 metres in the South East Zone.”

11.1 LOCATION AND SURVEYING

“Proposed drill sites were mostly located relative to existing and identifiable Russian drill collars, which were first identified from the Russian geological maps and then found and marked in the field in advance of drilling. With the large number of collars still visible in the field, there was much evidence of the overall reliability and accuracy of the Russian drillhole locations on maps and very little suspicion of possible mistakes on their part.”

11.1.1 Field Location

“The locations of all Russian and WNP drillholes to date are shown on Figure 8 [Figure 11.1].”

11.1.1.1 Russian drillholes

“To locate Russian drillholes in the field, their Gauss-Kruger coordinates were calculated on the Russian geology maps and converted into UTM. These UTM coordinates were then entered into portable GPS units to get close to the drillhole. In most cases, the drillhole collar is still visible to confirm the location. Due to the lack of precision of GPS, location with this method was, whenever possible, confirmed by triangulation from known points. In the case of an important Russian collar not being visible, its precise field location was requested from the surveyor.”

11.1.1.2 WNP drillholes

“Holes to be drilled were located in the field by triangulation from previously located Russian drillholes in areas where the Russian drilling density was sufficient to ensure a proper field location. In areas lacking the Russian collar references, the field location for the WNP hole to be drilled was generally determined by the surveyor.

“The hole to be drilled was marked on a wooden peg with number, direction, dip and estimated depth. One wooden peg was emplaced as a foresight in the drilling direction at about 15m from the dill-hole location. Two wooden pegs were emplaced as back sights on the back of the drilling line.”

11.1.2 Surveying

“The Gurvanbulag survey was generated using three Mongolian state geodetic base stations (#222, 009 & 102) belonging to the Bayundun Sum topographic grid and three additional points (#504, 014 & 307) from the Dornod Aimag geodetic grid. In the polygonometric grid, the “Khurten” point was used as the main hard point for location and height. The original coordinates in the Governmental records are in Gauss Kruger.

“Due to an insufficient number of base points for such a large area (30km from the SW Zone of Gurvanbulag to the east of Mardaingol), a professional geodesy company was hired to install 12 additional hard base points: 5 stations in the Gurvanbulag area, including 2 in the

SW Zone, and 7 points in the Mardaingol area. Besides the Khurten base point, the company used 3 more Mongolian state geodetic base stations, Ulaan for the Gurvanbulag area, Mardai and 1442 AS stations for the Mardaingol area.

“Topographic mapping was carried out using an electronic tachometric (Reflectorless Total Stations) PowerSet 2030, Sokkia, Japan, for field surveys. Data was processed on the SDR Map and Design program. This software enables translation of survey points into coordinates.

“Processing has been conducted in accordance with the book named “Location display mapping terminology and mark scale 1:5000, 1:2000, 1:1000 and 1:500” by the State Geodesy and Mapping Office of the Mongolian Government Implementation Agency, in 2001.”

Figure 11.1 Locations of Drill Holes

11.1.2.1 Surveying of drillhole collars

“All WNP drillhole locations entered in the database were surveyed after completion using the method described above. The surveyor recorded the collar of the drilled hole marked by a metallic or a PVC pipe. Visual checks of the relative positions of the various holes on the Gemcom maps were regularly done to avoid major mistakes, and comparison of the temporary handheld GPS coordinates with the survey data was also done as a quality control measure.”

11.1.2.2 Mapping

“Besides surveying drillholes, topographic mapping covered a limited area around the Gurvanbulag main shaft. This area includes part of block 1-06, the rock dumps around the shaft, the settling pond area and all former and new buildings in the fenced area of the mine site.”

11.2 DOWNHOLE SURVEYS

“Down-hole surveys were undertaken at 50m intervals from 0 to 300 metres, at 100 metre intervals below 300 metres and at the bottom of each drillhole. In the course of the program, an extra down-hole survey was systematically required as close as possible from surface to give a more reliable surface direction to the drillhole than the field geologist’s surface records.

“A Sperry Sun photographic instrument was initially used by one of the drill contractors (LandDrill), but was soon replaced with a Reflex single shot camera instrument, which automatically recorded the magnetic direction and the dip of the instrument. Table 7 [Table 11.2] tabulates drillhole deviations for the 2005 drilling program.

Table 11.2 Total Downhole Deviations in 2005 Drillholes

Drill-hole Deepest Azimuth Dip Drill-hole Deepest Azimuth Dip Number Test Deviation Deviation Number Test Deviation Deviation (m) (o) (o) (m) (o) (o) GCE5001 240 7.0* 0.5 GCE5002 320 2 0 GCE5003 300 7 1.7 GCE5004 331.2 3 0.5 GCE5005 None None None GCE5006 287 6.8 7.1 GCE5007 337 6.0* 3.1 GCE5008 282.5 9.4 5 GCE5009 279 7.3 5.3 GCE5010 269.9 8.4 7.3 GCE5011 250 7.6 5.3 GCE5012 120 0.4 0.6 GCE5013 250 6.1* 0.8 GCE5014 150 Vertical 0.7 GCE5015 50 Vertical 1.3 GCE5016 140 Vertical 1.9 GCE5017 140 Vertical 0.7 GCE5018 120 Vertical 1.7 GCE5019 130 Vertical 0.1 GCE5020 140 Vertical 0.3 GCE5021 140 Vertical 0.9 GCE5022 80 Vertical 0.1 GCE5023 60 Vertical 0.2 GCE5024 40 Vertical 0.7 GCE5025 60 Vertical 1 GCE5026 80 Vertical 1.1 GCE5027 90 Vertical 1.4 GCE5028 90 5.9 0.5 GCE5029 110 2.1 1.5 GCE5030 120 Vertical 1.9

GCE5031 130 Vertical 0.7 GCE5032 110 Vertical 0.3 GCE5033 75 0.2* 0.8 GCE5034 130 3.3 3.2 GCE5035 110 Vertical 1.1 GCE5036 50 4.3 1.5 GCE5037 110 0.9 0.8 GCE5038 130 2.7 3.4 GCE5039 51 Vertical 0.5 GCE5040 60 4.2 1.5 GCE5041 130 Vertical 0.7 GCE5042 160 Vertical 1.3 GCE5043 150 Vertical 0.9 GCE5044 150 Vertical 1.9 GCE5045 90 Vertical 0.3 GCE5046 59.75 Vertical 0.3 GCE5047 80 Vertical 0.5 GCE5048 129.7 0.8 0.4 GCE5049 129.85 Vertical 0.9 GCE5050 150 Vertical 0.5 GCE5051 150 Vertical 0.2 GCE5052 160 Vertical 0.3 GCE5053 180 8 1.1 GCE5054 500 11.6 0.7 GCE5055 501.5 15.4 1.8 GCE5056 271 7.4 0.2 GCE5057 519 26.7 0.4 GCE5058 500 23 1 GCE5059 540 35.3 1 GSW5060 200 5 3.4 GCE5061 50 2.2 1.6 GCE5062 110 5 2.5 GCE5063 120 3.1 0.7 GCE5064 75 0.5 0.1 GCE5065 139 1.3 0.6 GCE5066 100 0.3 0.4 GCE5067 99 1.5 0.9 GCE5068 239.55 4.8 0 GSW5069 200 7.8 1.5 GCE5070 170 4.4 1.2 GSW5071 465 11.6 6.8 GSW5072 400 7.4 3.3 GSW5073 250 2.5 1.1 GCE5074 462 18.5 2.7 GSW5075 178 5.2 1.3 GCE5076 170 4.3 0 GSW5077 210 4.8 0.1 GCE5078 170 4.5 0.3 MSW7001 150 1.3 0.9 MSW7002 150 3.9 1.6 MSW7003 61 3.1 0.4 MSW7004 133 4.2 1.5 MSE7005 300 8.9

“Every down-hole survey was recorded both on the field geologic log and in the survey database. Besides the records on the geologic logs, all the exposed films from the down-hole surveys and the field records from the drillers are filed with the daily reports from the drillers as a future reference.

“The deviation of the drillholes dip angles remains low at all depths while the deviation in direction steadily increases with depth for HQ core (to depths of ~300 metres). Below 300 metres, deviation dramatically increased in NQ core to an unacceptable level (35.3°). Drillers consequently changed the core barrel in December 2005 to use octagonal shaped barrels, which should reduce the deviation in deep holes for the 2006 program.”

11.3 CONSTRUCTION OF DRILLHOLE DATABASE FROM RUSSIAN DATA

“No information exists on the drilling procedures used by the Russians. All of the Russian data in the resource database has been digitized from hard copies of detailed plans and sections produced by the Russians. In many cases, the old drillhole collars have been identified in the field and surveyed, confirming the accuracy of their collar positions. The downhole traces have been meticulously digitized in 3 dimensions, under the supervision of Doug Blanchflower, who has described the process in the following section.”

11.3.1 Digitizing of Russian drillholes

“Between 1985 and 1990 Russian personnel compiled comprehensive annual reports documenting the mineral exploration activities on their properties, including resource estimations. A number of very detailed plans and sections accompanied these reports showing the locations, traces and geological and geochemical results from various surface and underground drilling programs. In April 2005, WNP produced scanned raster images of these plans and sections, the only readily available data documenting the extensive drilling work undertaken on the properties.

“Since May 2005 WNP personnel have been using the plan and section raster images to re- create drilling databases for work conducted within the Gurvanbulag Central (designated ‘GCE’) and Gurvanbulag Southwest (designated ‘GSW’) mineralized zones. These zones were of immediate interest and they are relatively well documented in the Russian exploration reports.

“Topographic plans for the GCE and GSW zones were created by digitizing sections of available 1:2000 and 1:10,000 topographic and drillhole location maps. All surface drillhole collars within the digitized areas were located, marked and their northerly and easterly Gauss Kruger coordinates recorded. In addition, the collar elevation for each plotted hole was estimated for later comparison with those measured from available vertical sections.

“Drill holes were commonly plotted by Russian personnel on vertical sections at a scale of 1:1000, 50 metres apart and oriented at 310-130°. Each section has both vertical and plan plots of the surface and, in many cases, the underground drill holes within the section length. In areas of intense underground drilling, detailed vertical sections were plotted at a scale of 1:500, at 25-metre intervals, and oriented usually at 310-130°. WNP personnel have used CAD software to display, scale and finely orient the various raster images; after which all downhole surveys, geological and assay data for each verifiable drill hole were recorded on a ‘matrix’ style form and later entered into a computerized spreadsheet-style database. The exact collar elevation for each drill hole was recorded from vertical section plots and compared to the one estimated from topographic and drill hole plans. More than 90% of the collar positions have been confirmed by recent surveys. Downhole survey information was collected by measuring azimuths and dips of each drillhole trace at each inflection point along its length. Intercepts of lithological units and their documented alteration and mineralization were recorded along each drillhole trace. In addition, mineralized intercepts with documented percent (%) uranium and interval lengths were recorded relative to distances from their drillhole collar.

“Approximately 2,300 Russian surface and underground drill holes tested the GCE zone and well over 100 surface holes tested the GSW zone. However, many of these holes were either not documented, were documented but their sections are not available, or they were plotted at extreme edges of vertical sections where important information was missing. Only reliable and verifiable drill hole data were collected; thus, no data were collected where either the collar and/or upper sections of drill holes were missing (i.e. often where the section title block obscured the plan strip), or where the lower portion of the drill hole ran off section. Upon conclusion of the drill hole data collection, drilling results for 65 surface holes within the GSW zone and 806 surface holes within the GCE zone were entered into their respective

drilling databases. In addition, drilling results for 692 underground drill holes have also been compiled for the central portion of the GCE zone.”

11.4 DRILLING RESULTS

“The 2005 drill program aimed at: • testing the near-surface Russian C1 category resource block 1-06 in the Gurvanbulag Central Zone to allow a geostatistical comparison of grades, thicknesses and distribution of the uranium mineralization with the Russian work; • verifying other mineralized zones in the Gurvanbulag Central Zone; • exploring for new mineralization within or near the Gurvanbulag Central Zone; • exploring for new mineralization near the Gurvanbulag Southwest Zone; • exploring for new mineralization in the Mardaingol area.”

11.4.1 Resource Block 1-06

“Dr. Gerald Harper recommended a confirmation surface drilling program targeting the Russian resources block # 1-06. All mineralization in block 1-06 is hosted in the Gurvanbulag main horizon, a structure conformable to bedding, which dips gently to the southeast. Resources in block 1-06 in the C1 category as quoted in Trikilov & al. (1986) are 362,206 tonnes grading 0.261% U (or 0.31% U3O8).

“Drilling of this block presented two main advantages: • a higher than average grade for surface holes, • this is the only block defined in the C1 category close to surface, allowing relatively short holes.

“The program was designed to “cover the block adequately to allow an independent resource calculation to be made to compare with the Russian calculation and to allow combination of both data sets to provide for a higher reliability resource calculation”. Initially, forty drillholes were planned to test block 1-06. Two holes in the series experienced technical problems and had to be re-drilled. GCE5061 and GCE5062 were re-drilled as GCE5036, and GCE5063 was re-drilled as GCE5039. Two additional drillholes, GCE5070 and GCE5076, were later drilled to increase the size of the mineralized zone. Drillholes were positioned at 50-metre centres and numbered GCE5014 to GCE5053. Spacing also included some very closely spaced holes (17m spacing between GCE5015, GCE5016, GCE5017 and GCE5020) to document evidence for an erratic distribution of mineralization. The totals for the program reached 3,441 metres in 45 drillholes; assay results are given in Table 8 [Table 11.3].

Table 11.3 WNP Drilling Results from Russian Resource Block 1-06 (Reported intervals are >94% of true width)

Drill-hole Section Intercept Interval U3O8 (m) (m) [%] GCE5014 9400N 113.1-115.1 2 0.09 GCE5015 9400N 99.0-101.1 2.1 0.46 GCE5016 9417N 105.4-106.8 1.4 0.16 GCE5017 9433N 110.0-112.1 2.1 0.37

GCE5018 9450N 96.4-98.5 2.1 0.28 GCE5019 9450N 103.7-105.5 1.8 0.19 GCE5020 9450N 115.2-117.0 1.8 0.09 GCE5021 9450N 118.7-120.5 1.8 0.17 GCE5022 9450N 70.3-71.1 0.8 0.05 GCE5023 9450N - - - GCE5024 9400N - - - GCE5025 9400N - - - GCE5026 9400N 60.7-63.3 2.6 0.12 GCE5027 9400N 70.8-73.0 2.2 0.09 GCE5028 9400N - - - GCE5029 9400N 83.0-84.2 1.2 0.05 GCE5030 9350N 102.4-104.4 2 0.42 GCE5031 9300N 96.1-98.9 2.8 0.18 GCE5032 9300N - - - GCE5033 9300N 62.0-63.8 1.8 0.23 GCE5034 9300N 106.5-107.3 0.8 0.06 GCE5035 9300N - - - GCE5036 9300N Re-drilled as GCE5061 and GCE5062 GCE5037 9300N - - - GCE5039 9400N 101.9-102.9 1 0.12 GCE5040 9500N 144.7-146.0 1.3 0.12 GCE5041 9500N 118.4-119.4 1 0.12 GCE5042 9500N 111.0-113.0 2 0.23 GCE5043 9500N 121.0-123.1 2.1 0.18 134.8-137.1 2.3 0.1 GCE5044 9500N - - - GCE5045 9500N - - - GCE5046 9500N - - - GCE5047 9550N - - - GCE5048 9550N 111.0-111.9 0.9 0.1 GCE5049 9550N - - - GCE5050 9550N 123.6-130.1 6.5 0.44 GCE5051 9550N 133.2-137.6 4.4 0.37 GCE5052 9550N 147.7-150.0 2.3 0.16 GCE5053 9550N 158.6-161.0 2.4 0.21 9300N 169.9-172.0 2.1 0.18 GCE5061 9300N Re-drilled as GCR5062 GCE5062 9300N 88.1-90.0 1.9 0.72 GCE5063 9400N 102.7-103.2 0.5 0.07 GCE5070 9600N - - - GCE5076 9600N 147.0-149.7 2.7 0.14

“Mineralised intersections belong to the main Gurvanbulag horizon. Drilling indicates a strong positive correlation between uranium mineralization and the intensity of fracturing and alteration. Using comparable parameters with Russian composites used in their resources calculations, grades of the WNP intersections compare well with the corresponding Russian drillhole results in the block while the thicknesses of the WNP intersections tend to be slightly greater. However, a full geostatistical comparison is required to fully evaluate the two sets of results (see Section 16.5.4) [of SRK (2006)]. The program also confirmed the erratic distribution of the mineralization as shown by the results of the 17m spaced drillholes with

grades of 0.46 % U3O8 in GCE5015, 0.16 in GCE5016, 0.37 in GCE5017 and 0.09 in GCE5020. Widths for these intersections are given in the table below [see Table 11.4].”

11.4.2 Verifications in Gurvanbulag Central Zone

“A program of 13 drillholes was initially planned to verify continuity and grades of the mineralization in the Central Zone of the Gurvanbulag deposit. It was completed with another set of 4 drillholes: GCE5074, GCE5075, GCE5077 and GCE5078. All these drillholes reached target except for GCE5075 which had to be abandoned due to loss of water when it drilled into the Russian hole it was supposed to twin.

“As for block 1-06, grades of the intersections favourably compare with the Russian drillhole results in the block. The following drillholes were twins of Russian drillholes: • GCE5001 twinned Russian drillhole 1341; • GCE5074 twinned Russian drillhole 1380; • GCE5075 designed to twin drillhole 4884, but did not reach target; • GCE5077 twinned Russian drillhole 4862; • GCE5088 twinned Russian drillhole 4617.

“Most drillholes of this program combined testing of the main Gurvanbulag horizon with testing of uranium-bearing vertical structures higher up in the stratigraphy.

“However, most significant mineralized intersections came from the Gurvanbulag horizon. Drilling vertical structures led to only two uraniferous intersections, one in GCE5005 and the other in GCE5012.

“The comparison of twins between the Russian and the WNP drill results (Table 9) [Table 11.4] shows that WNP twins of Russian generally returned lower results. The most likely explanation combines two factors: • The distribution of the uranium mineralization is erratic throughout the deposit. • The WNP twins aimed at duplicating some of the highest grade intersections. With an erratic distribution of the uranium mineralization, the odds of reproducing a similarly high value are much reduced.”

Table 11.4 Comparison of Results from WNP and Russian Drillholes

WNP Drill- Section Interval U assay Russian Interval U assay hole (m) (%) Drill-hole (m) (%) GCE5001 9600E 2.2 0.18 1,341 2.4 0.981 GCE5074 10450N 4.3 0.62 1,380 2.9 0.289 0.7 0.51 9.7 0.359 GCE5075 10550N Did not reach target 4,884 0.7 0.344 4.5 0.014 GCE5077 11100N 0.8 0.05 4,862 2.7 0.018

GCE5079 11200N - - 4,617 5.3 0.114

11.4.3 Exploration near Gurvanbulag Central Zone

“Eleven drillholes were dedicated to exploration in the immediate vicinity of the Central Zone of the Gurvanbulag deposit. Drillholes GCE5054 to GCE5059 were designed to test the main Gurvanbulag mineralized zone at depth east of the main shaft. Drillholes GCE5064 to GCE5067 tested the same horizon closer to surface down dip from the surface trenches of line 9 in grid GA1. Drillhole GCE5068 tested vertical fracture controlled mineralization below a mineralized Russian trench. Table 10 [Table 11.5] presents the results of these drillholes. Mineralization was intersected in 4 of the 11 holes, but grades are low.

Table 11.5 Results of Exploration Drillholes in the Gurvanbulag Area

Hole ID Section Intercept Interval U3O8 Assay (m) (m) (%) GCE5054 10400N - - - GCE5055 10400N - - - GCE5056 10400N - - - GCE5057 10600N - - - GCE5058 10600N - - - GCE5059 10600N 491.0-492.0 1.0 0.09 GCE5064 11400N - - - GCE5065 11400N 119,6-121.6 2.0 0.07 GCE5066 11500N - - - GCE5067 11600N 37.5-41.5 4.0 0.06 GCE5068 10350N 224.0-226.0 2.0 0.08

“The results of this test program suggest that in view of the density of prior Russian drilling, it will be difficult to find undiscovered attractive mineralization in the immediate vicinity of the Russian resource blocks. Exploration drilling aimed at adding resources should rather be concentrated in un-drilled or weakly drilled areas such as the GB1 target.”

11.5 CONCLUSIONS OF WNP 2005 DRILL PROGRAM

“The 2005 drill program: • confirmed overall grades and thicknesses of Russian exploration work for the Gurvanbulag U mineralization; • confirmed the erratic distribution of mineralization, both with close spaced drilling in block 1-06 and by twinning Russian drillholes; • confirmed that a geostatistical approach is necessary to convert Russian resources to NI 43- 101 compliant resources; • intersected very high grade uranium mineralization, confirming therefore the reality of higher grade mineralized zones; • confirmed that the potential for increasing resources lies with drilling undrilled or low- density drilled areas on the property. In total, the drillhole database contains data for 906 surface drillholes (including 110 WNP holes) and 692 underground drillholes.”

As noted above, the resource estimate for the Central Zone of the Gurvanbulag deposit completed by SRK (2006) was based on drill data to March 18, 2006.

12.0 SAMPLING METHOD AND APPROACH

The sampling method and approach is described in SRK (2006), as follows:

12.1 DATA ACQUISITION ON CORE

“The text from this section has been adapted from WNP’s geological field manual dated April 2005. Geological logging forms were designed between April 2005 and the start of drilling in June. Operational modifications were implemented during the course of the program.”

12.1.1 Core Handling

“Every core box was labelled at the drill site with its hole identification number and box number clearly marked with water-proof black felt marker on top and front end of each core box. The drillers used metric labelling to two decimal places on their depth blocks. Re-usable plywood lids were used to close the wooden core boxes before transportation. The lids were securely fastened as each core box was filled. All core boxes were gently loaded and unloaded to avoid tipping and movement of the core within the boxes and core boxes were properly secured within vehicles.

“A separate outside vehicle-accessible staging area was set up for each drill at the entrance of the core logging building. Core boxes were unloaded by the drillers from the transport vehicle at the end of each shift. Technicians arranged the core boxes at the beginning of the shift in chronologic order, removed the lids and generally washed any drilling fluid or mud from the boxes while in the staging area. Core boxes were then carefully carried to the core logging tables within the logging facility. The core boxes were arranged chronologically on the core logging table.”

12.1.2 Depth and Recovery Measurements

“Once the core boxes were arranged chronologically on the core logging table, the geotechnical assistant marked core at one metre interval on all core boxes. Core recovery was recorded by the geotechnical assistant immediately after the core boxes arrived in the core logging facility. The lengths and percentage of core recovered was recorded for each drill run in the “Geotechnical Drill Log”. The overall sum of the measured intervals divided by the total length of the drillhole gave the core recovery for one drillhole. The first few metres of overburden were not considered in the recovery calculations at Saddle Hills and core recoveries expressed for the project only reflect the recoveries of rock. Average and minimum core recoveries for each hole are presented in Table 12 [Table 12.1].

“Lastly, after the core boxes had been marked with their respective depths, a 10- by 4- cm aluminum tag was permanently embossed with the drillhole identification, core box number, date and depths of drill core inside the box. The tag was then securely stapled to the front end of the core box as a permanent, weather-proof label.”

Table 12.1 Core Recoveries for 2005 Drill Program

Drill-hole Average Minimum Drill-hole Average Minimum Number Recovery Recovery in Number Recovery Recovery in (%) a Drill Run (%) a Drill Run (%) (%) GCE5001 97.2 47.8 GCE5002 97 12 GCE5003 93.8 62.3 GCE5004 95.9 46.6 GCE5005 97.2 85.2 GCE5006 98.5 91.8 GCE5007 99.1 88.5 GCE5008 97.5 64.3 GCE5009 97.4 24.6 GCE5010 99.9 95.1 GCE5011 99.6 97.7 GCE5012 99.2 95.7 GCE5013 95.6 14 GCE5014 89.5 46.7 GCE5015 95.5 80 GCE5016 96.8 74 GCE5017 97.7 78.3 GCE5018 97.8 90 GCE5019 94.2 21.4 GCE5020 97.3 69.2 GCE5021 98.4 81 GCE5022 95.8 82.7 GCE5023 96.5 75 GCE5024 93.8 40 GCE5025 96.1 85 GCE5026 95.3 68.9 GCE5027 96.9 66.7 GCE5028 97.4 50 GCE5029 97.4 64 GCE5030 97.5 81.5 GCE5031 97.4 81 GCE5032 96.9 66.7 GCE5033 96.8 73.1 GCE5034 98.4 83.5 GCE5035 93.2 44.4 GCE5036 98.2 93.3 GCE5037 98.8 85 GCE5038 98.2 86.4 GCE5039 99.1 94.7 GCE5040 96.3 24 GCE5041 99.5 96.7 GCE5042 99.2 91.8 GCE5043 99.3 85.7 GCE5044 97.4 33.3 GCE5045 99.4 95.1 GCE5046 99.5 92.6 GCE5047 99.2 91.8 GCE5048 99.6 96.7 GCE5049 93.7 78.7 GCE5050 95.5 85.2 GCE5051 95.5 90.2 GCE5052 94.3 90.2 GCE5053 95.8 90.2 GCE5054 96.4 62.3 GCE5055 96 55 GCE5056 96.7 79.7 GCE5057 96.7 15 GCE5058 97.3 63.5 GCE5059 97.2 85 GSW5060 96.8 90 GCE5061 99.4 96.7 GCE5062 98.1 89.3 GCE5063 96.4 85 GCE5064 96.2 90.2 GCE5065 96.8 68.6 GCE5066 94 82.1 GCE5067 94.6 55.8 GCE5068 97.3 90.2 GSW5069 97.3 75 GCE5070 98.7 81.3 GSW5071 98 76 GSW5072 97.4 67.7 GSW5073 96 72.7 GCE5074 97 83.3 GSW5075 96.4 90 GCE5076 98.6 96.3 GSW5077 97.7 91.7 GCE5078 94.4 59.3 MSW7001 96.7 74.3 MSW7002 97.5 46.7 MSW7003 99 74.6 MSW7004 97.8 86.1 MSE7005 90.9 4.3

12.1.3 Quick Log and Radiometric Scan

“While the assistant was re-labelling the depth blocks and doing the core recovery measurements, the field geologist quickly examined the drill core for notable geologic features, such as mineralization, stratigraphic or intrusive contacts, alteration changes, etc., and prepared a quick log outlining the main units and features on the drillhole for the day. The geologist also scanned the core to identify any radioactive zones of interest. Results of this radiometric scan were presented on the quick log that also included core recoveries in the various geological units.”

12.1.4 Geotechnical Logging

12.1.4.1 Rock Quality Designation

“Rock quality designation (RQD) was recorded for each drill run. Drilling induced breaks and breaks from loading core boxes were not included. Veined core pieces were lightly hammer tapped. Those pieces that remained intact were included as the solid fraction of core for the RQD determination.”

12.1.4.2 Discontinuities

“The number of discontinuities (broken or open joints, shears, veins, etc.) per drilled interval was recorded, not including any breakage due to drilling or loading core boxes. It is uncertain whether or not the early records, collected while the staff were being trained, include healed joints or veining that should have been excluded.

“The following structural features and characteristics are included in the geotechnical logs:

1. Joint 2. Bedding 3. Fault 4. Shear 5. Schistocity 6. Cleavage 7. Foliation 8. Vein 9. Contact

“The average shape of the joints was noted using the descriptors: 1. stepped 2. undulating 3. planar

“The average roughness of the joints was noted using the descriptors: 1. rough 2. smooth 3. slickensided”

12.1.4.3 Hardness or Rock Strength

“Rock hardness, based on the use of either a pocket knife and/or geological hammer, was recorded using codes from “0”, being extremely weak rock, to “6”. The tables used as reference for the Saddle Hills work were taken from Knight Piesold’s “Geotechnical data collection for exploration geologists” manual.”

12.1.4.4 Weathering

“Rock material weathering is the degree of weathering of solid rock pieces between fractures and varies from fresh to extremely weathered. The average weathering conditions of a drilling interval was recorded on the “Geotechnical Drill Log” form according to a table from Knight & Piesold’s “Geotechnical data collection for exploration geologists” manual.”

12.1.5 Radiometric Scan Log

“The geotechnicians or the geologist did a systematic radiometric scan of the drillhole and recorded the minimum and maximum total count values for each interval of 1m from the start to the end of the drillhole. These values were reported, along with an average of the two values, in the “Radiometric Log” forms, and the averages were later transferred to the radiometric table of the database. Graphs of the averaged radiometric values were made to help the logger in the identification of zones to be sampled and project management for easy verifications.

“In 2005, most measurements were made using two old McPhar TV1A instruments. These were difficult to calibrate and did not hold calibration too well. Therefore, this table of the database gives qualitative measurements on a hole by hole basis rather than accurate quantitative measurements.”

12.1.6 Core Logging

“The matrix logging form used for the 2005 drilling program resulted from the combined logging experience of project managers and advisors. It documents the geological logging results in a form that can be readily input into the Gemcom software program.

“The top part of the log’s first page contains all necessary drill hole information, including: an area for a small sketch map to show the drill hole location relative to other survey monuments or existing surveyed drill hole collars; drill hole location coordinates; total drilled length; casing length; core size; reduction depth if required; dates of drilling, logging and data input; names of the logging and data input personnel; and whether the casing was removed or left in place. In addition, there is a section in the header area for recording the collar azimuth and dip, and subsequent down-hole survey measurements with the type of down-hole test.

“The following features were logged in each of the holes: • interval, • rock coding, • sampling, • mineralization, • geology, • alteration,

• structure, • comments and descriptions.”

12.1.6.1 Interval

“Matrix logs require that all described geological intervals be primary intervals. In other words, all described intervals must have a “from” and “to” drilling length recorded without any secondary or “nested” descriptions.

“To allow for quantitative estimations to be meaningful, several consecutive lines may have the same rock code in a single geologic unit, differences between the lines being either in sampling, mineralization, geology, alteration or structure estimates or codes.”

12.1.6.2 Rock Coding (Lithology)

“Major and minor lithological units are described on the matrix geologic log with a four-letter abbreviation in capital letters. A major lithological unit is the dominant rock type within the stated drilling interval while the minor lithological unit is the subordinate rock type. All geologic descriptions will have a major rock type, but minor rock types are commonly used to record textures.

“Lithological unit names are commonly abbreviated with the first letter, sometimes the first two letters, describing the general type of rock. Intrusive rocks are stated by the letter “I”, flows by the letter “F”, sediments by the letter “S”, volcano-sediments or tuffs by the letters “TF” and metamorphic or metasomatic rocks by the letter “M” (Table 13 [Table 12.2]). The other two or three letters summarize the rock name in a way it may easily be memorized. For example, a gritt is “SGRI” while a tuff gritt is “TFGR”.

Table 12.2 Lithology Codes for the Gurvanbulag 2005 Drill Program

Intrusive and Sub-Volcanic Volcanic Rocks Sedimentary Rocks Volcano-sedimentary Rocks Rocks IDIO Diorite FBAS Basalt SARE Arenite TFAG Agglomerate IQZD Quartz diorite FAND Andesite SCNG Conglomerate TFAN Andesitic tuff IGRA Granite FDAC Dacite SGRI Gritt TFDC Dacitic tuff (Undivided) IGRD Granodiorite FELS Felsite SLMS Limestone TFQF Quartz- feldspart tuff IMNZ Monzonite FLAT Latite SMDS Mudstone TFQZ Quartz-eye tuff IMND Monzodiorite FRHY Rhyolite SDST Sandstone TFTR Tracytic tuff IQZM Quartz monzonite FOBS Obsidian SLST Siltstone TFXX Tuff (Undivided) IFPO Plagioclase FJAS Jasperoid SHAL Shale TFVT Vitric tuff porphyry IQFP Quartz-feldspar TFGN Ignimbrite porphyry IPOR Porphyry Metamorphic/Metasomatic Rocks TFEL Felsite (Unidivided) (rhyolite tuff) ISYE Syenite MGNS Gneiss TFGR Tuff gritt

MARB Marble TFST Tuff sandstone MSCH Schist TFLA Lapilli tuff (Undivided) MTCL Tectonic clay TFSL Tuff siltstone TFCG Tuff conglomerate Sub-rock Type Codes PE Perlite CL Coal, carbon PL Pillowed DV Devitrified QZ Quartz-eye HY Hyaloclastic FL Fluidal BR Breccia PO Porphyritic SP Spheroidal CO Columnar MA Massive LI Lithic VE Vesicular BA Banded Miscellaneous OVBN Overburden CONT Contaminated NOCR No core recovery

“These logs are far more detailed than the lithological information provided on the Russian sections, and the geological model produced for this study lumped all lithologies into three units: FOBS, TFEL (HW) and TFGN (FW).”

12.1.6.3 Sampling

“The sampling section of the log contains 4 columns: • sample number, always starting with A0, • radiometric count of the sample in counts per minute, • XRF results in ppm U; • assay results in ppm U.

“The sample number is the only column of this section filled by the geologist while logging. The radiometric count for the sample was initially recorded, measuring the various channels of the sample bags over more than 1 minute over 5 different faces of the sample. As this operation proved to be too time consuming (over 15 minutes per sample), and the XRF result provided a rapid estimate of the sample’s assay, it was abandoned after a month. Both XRF and assay result are reported on the spreadsheet when available, after the logging.”

12.1.6.4 Mineralization

“The mineralization section contains several columns with the most common economic and non-economic minerals. Within each column the geologist would write in the estimated percentage of each mineral. The various columns of the “Mineralization” section are: • pitchblende – uraninite (%) • molybdenite (%) • pyrite (%); • galena (%); • uranophane (%); • uranium mineral 1 (%) • other

“As pitchblende and uraninite may not be distinguished, even with a magnifying lens, they are quantified together. The last two columns are left for other minerals, the recording of which may prove useful in the future.

“In 2005, quantitative estimates by WNP’s rookie staff proved unreliable. For instance, estimates of 30% pitchblende were made for GCE5003 from 314.35 to 314.8m. The assay result only returned 1.18% U3O8. Despite progress, an estimate of 2% uranophane in GCE 5076, drilled in December, resulted in an assay of 0.27%U3O8.

“Further effort towards reasonable estimations will be made in 2006.”

12.1.6.5 Geology

“Several features of the rocks may be described in a qualitative or quantitative way. This section details: • oxidation level of the rock (oxidized, fresh reduced), • shade and color of a rock with 1 letter for the first and 2 for the latter, for instance a rock may be medium brown (M BR), dark grey (D GY) or light green (L GR), • sedimentary features and orientation, • maximum size fragment in millimetres, • fragment / matrix ratio, • grain size in millimetres.

“The relevant codes for this section of the logs are listed in Table 14 [Table 12.3].”

Table 12.3 Geology, Structural and Alteration Codes for 2005 Drill Logs

Oxidation code Colour Code Colour Code (cont.) Sedimentary Features Symbol Description 1st 2 Description 3rd letter Description Symbol Description letters R Reduced BK Black D Dark UN Unconformity O Oxidized GY Grey M Medium GB Graded bedding F Fresh BR Brown L Light BD Bedding PU Purple RM Ripple marks BL Blue GR Green YE Yellow OR Orange RE Red WH White Alteration Intensity Codes Structural Codes Structural Codes (cont.) Symbol Description Symbol Description Symbol Description 6 Intense FB Fault breccia 6 <0.5 cm apart 5 Strong FN Normal fault 5 0.5-1 cm apart 4 Moderate FR Reverse fault 4 1 cm – 10 cm apart 3 Weak FS Strike-slip fault 3 10 cm – 50 cm apart 2 Very weak FL Fault (Undivided) 2 50 cm – 1 m apart 1 Trace SH Shear (Undivided) 1 1-3 m apart 0 No alteration BR Broken zone 0 None

12.1.6.6 Alteration

“The alteration section contains several columns with the most common alteration minerals. Within each column the geologist enters a semi-quantitative estimate of the intensity of the alteration by this mineral (Table 14) [Table 12.3].

“The semi-quantitative approach was preferred to a true quantitative approach as it is very difficult to estimate true mineral percentages in case of fine grained pervasive alteration. The various minerals of the “Alteration” section are: • quartz • carbonate • kaolin • hematite • limonite • manganese • fluorite • chlorite • sericite • other,

“The column “other” was changed to “unknown soft” (mineral) during the course of the program. This mineral still has to be identified.

“A board with the various levels of alteration for each of the minerals in each type of rocks is being prepared. It was not completed in 2005 and there still are discrepancies between loggers.”

12.1.6.7 Structure

“The three columns dedicated to structure in the geologic log are sufficient to describe the dominant structures in the interval: • the “Structure type” column records the type of structure using a 2-letter code such as FL for fault and SH for shear, • the “Spacing” column records spacing between these structures using numbers from 0 (none) to 6 (<0.5cm apart); • the “Orientation” column records the angle of the structural feature to the core axis.”

12.1.7 Specific Gravity Measurements

“During the 2005 drilling program, specific gravity (“SG”) measurements were recorded every 10 metres. Measurements were at even depth at 10, 20, 30m etc., but the exact sample around each depth was chosen by the field geologist in view of the competence of the core.

“In the first part of the program, a rather imprecise method was used and all SG measures collected by this method were discarded from the database. After October 19, 2005, the mistakes were identified and the following method was effectively used.

“The sample was weighed in air and in water. The water measurement was problematic as the plate and suspension beams of the balance were also submerged and contributed to the displacement of the water. This problem was partially avoided by subtracting the weight of the balance apparatus in water from the measured weight of the sample in water. However, the higher level of water in the bucket, upon introduction of the sample, resulted in a minor imprecision due to the resulting submersion of a greater length of the rods of the balance pan at this stage.

“Considerable effort was made to validate this method, using comparative tests with suspension of the sample by a thread, rather than the balance pan, and by decanting volumes of water equivalent to the volume of the samples being measured, and WNP is confident that their measurements are valid to within +/-1%. SRK has directly observed the density measurements in progress, as well as, reviewed the large resulting database, and is satisfied that they are adequate for resource estimation.

“Note that the densities are referred to here as ‘specific gravity’; however, they are, in fact, bulk density measurements of intact core pieces. Since the porosity of these samples is negligible, the reported densities are not expected to differ measurably from the actual specific gravities. These data were used directly in the resource model to estimate the density of the mineralized zones, with no consideration of fracture frequency and/or other features that could contribute to a lower overall bulk density of volumes larger than a single piece of core.”

12.1.8 Photography of Core

“After geotechnical and geological logging, the entire core was photographed in detail prior to any core splitting and sampling. Two core racks were constructed to support 3 boxes of HQ or NQ core. These core racks are approximately 1.5 m high by 1.5 m wide with a 45° sloping front. There is a permanently affixed metric rule between the three rows of core boxes which provides a scale for the photographs.

“Digital colour photographs were taken in good light, perpendicular to the core boxes. In this manner the entire lengths of all three core boxes could be photographed without shadows. The digital images were downloaded daily to an on-site computer. Selection was made between several pictures of the best one for each series of core boxes. The pictures were then filed with the logs. The rejects were not discarded but also filed in another folder in case of future need.”

12.1.9 Core Storage

“Following all logging and sampling operations, core boxes were transported directly to a core 2 storage area. The area is located outside, on a flat, 800 m area. The 16,500 metres of core from the 2005 drill program have been stored in piles of boxes stacked to a height of 1.80 m with lids only on the top boxes. Total capacity of the storage area is about 73,000 metres.”

12.2 DRILL CORE SAMPLING

“After a spectrometer scan of the core, the geologist in charge of logging a drillhole chooses the sampling interval, taking into account the radiometric results and the geology. It was decided that samples must be composed of one geological unit. For mineralized zones of 2 metres and

less, sample lengths should be 0.5 metres and for zones over 2 metres, sample lengths should be 1 metre. Two to three samples are to be collected on either side of the mineralized intersection.

“WNP had some difficulty in enforcing this protocol on their predominantly Mongolian geologist staff. Early sampling intervals did not always match geological boundaries, and at year end (Dec 2005) samples with a length over 1 metre were still being recorded in mineralized zones despite frequent reminders to the geologists.

“The radiometric scans assisted in choosing the sampling zones and as a later check for inadvertently forgotten sampling zones. This check is done prior to the shipment of samples to the Laboratory. At the beginning of the program, prior to the systematic implementation of the graphs, sampling of a forgotten zone was often late, and usually these samples became part of a later shipment. Later in the year, re-sampling occurred as shown by disrupted sample number sequences in the drillhole database, but all samples were sent in the same sample shipment.

“Sampling occurs after the core has been photographed and the different geotechnical measurements of the core have been made. The geologist marks the sample length on the core, and then the technician splits the core piece by piece into two halves. One half of the core is returned to the core box for storage and the other half is placed in a plastic sample bag. A pre- numbered tag, corresponding to the sample number in the geological log and to a sampling book form, is added to the bag immediately after the split. The sample number is also marked on the bag with a felt pen.

“Initially, a different procedure was used. The geologist would give the helper a handwritten listing of as many as 15 numbers on pieces of paper. The helper would then insert these into as many bags in advance of the sampling. The geologist only added tags later. This method, used by Russian trained geologists, presents a high level of risk for mistakes. During this procedure, all sample bags stay open in the working room close to the core splitter. Besides a risk of inexperienced staff making numbering mistakes, the method also runs the risk of contamination by flying chips coming from the manual core splitter.

“To avoid the risk of wrong numbering, the tag coming directly from the sample book is placed immediately into the plastic bag. The sample bag is then moved quickly to a safe storage room before shipment.

“The appropriate procedure would be to immediately seal the sample bags after the introduction of the tag. However, to reduce drying time at the laboratory, samples are dried at room temperature per special request from WNP. Thus, the bags are not immediately sealed to allow the samples to partly dry while stored before shipment. The open bags are kept in that closed room until shipment to Ulaanbaatar. At the end of the year, the Alex Stewart preparation laboratory was authorized to use a low temperature oven to dry samples at temperature below 40°C. But the procedure in the field was not changed to reduce any risks arising from disturbing the routine of the technicians.”

12.3 SHIPMENT OF SAMPLES

“Just before loading samples in the WNP truck, sample bags are closed and set in sequential order in the core shed for checking. The sample listing is prepared daily from drillhole “sample books” by the geologist in charge of the drillhole (or the trench). For confidentiality reasons, no drillhole numbers or depths of the samples are reported on these listings.

“Next, verification is done of the samples sorted in the core shed with the shipment list. Each sample number is checked against the list by a technician and the QA/QC geologist responsible for the shipment. Tags for standards, blanks, PD2, CD2 are collected from the sample books of the different geologists. If there is a difference between the shipment list and the samples, sample books are checked to find the difference. At the beginning of the program, many differences were found, most of them due to forgotten field duplicate preparation or to samples stored in the wrong place. Discrepancies between samples lined up and listings were rare during the second half of the 2005 drilling program, except for QA/QC tags which are still often forgotten inside the sample books.

“When sample shipment lists and the sample bags, plus QA/QC tags are in agreement, samples bags are loaded in a wooden box in the truck. A special box with metal and wooden reinforcements is used to carry highly radioactive samples. Radiation inside the truck with this special box is measured prior to allowing the transportation to go ahead. In 2005, only one series of samples was placed in the special box and sent as a special shipment: SH-2005-09. Radiation checks have indicated that the levels in the truck are similar to those of the core logging room background.

“The person in charge of the shipment has to count the sample bags being loaded and double check his count with the total number of samples in the shipment list. Boxes are then closed and the truck can leave the site with envelopes containing letters and listings, and proceed to move the samples to the preparation laboratories of Alex Stewart and WNP’s XRF laboratories in Ulaanbaatar, and ultimately for shipment to Actlab in Canada.

“Normal transportation by road takes about 1 or 2 days. Upon arrival of the truck, the Alex Stewart laboratory confirms this. Initially, this was done through the XRF technician, but later it was done directly by mail from Alex Stewart. Above two days of transportation, explanations are systematically required. Twice, transportation took 7 days. The old truck used then has been changed since.

“The WNP technician had been instructed to send the letter to Actlab in Canada with the samples. However, as Actlab reported receiving most shipments without any instructions, the procedure had to be changed. In the second part of the drill program, all letters and listings were sent by email from the camp to the laboratories.

“Complete shipment information with the corresponding drillhole and sample depths are sent by email to management officials Gerald Harper and Wayne Roberts. In this form, the QA/QC samples are clearly indicated.

“All exploration samples sent from the Saddle Hills project are assigned a shipment number. They include regular samples to be assayed, rocks sent for thin sections, or round robins of home-made standards. For an immediate request of re-assay, the initial shipment number is used. For later control, a new shipment number is used for samples leaving the camp or rejects leaving their storage in Canada. The same procedure will be used for metallurgical or geotechnical samples as it helps tracking them.

“Shipments numbers for 2005 are SH-2005-01 to SH-2005-30. Most of these shipments were sent to Actlab in Canada, a few to Becquerel, also in Canada, and only two for thin sections and/or special analysis.”

13.0 SAMPLE PREPARATION, ANALYSES AND SECURITY

Sample preparation, analyses and security are described in SRK (2006), as follows:

13.1 SAMPLE PREPARATION AND ANALYSES

13.1.1 Alex Stewart Laboratory, Ulaanbaatar

“In 2005, all samples were prepared at the Alex Stewart Laboratory in Ulaanbaatar. The laboratory received the sample shipments directly from the Saddle Hills camp. Sample preparation at the laboratory included the following steps: • Drying. For the first shipments, the laboratory was not allowed to use any oven to dry the samples for fear of isotope changes due to heat. “Accordingly, the delay in preparation was sometimes quite long, especially with clay-rich samples. As described earlier, in an attempt to reduce drying time, the sample bags are left open at the mine site in a safe room, but this is neither desirable nor efficient. So, for the last shipments of the year, Alex Stewart was allowed to use a low-temperature oven to dry samples at less than 40°. (SRK has not reviewed the relevant literature describing the effects of heat on uranium sample grades, but defers here to the preferences of WNP management. Significant differences in grade are not expected between WNP’s different drying methods). • Crushing to <2 millimetres. • Splitting 500g for a normal sample, twice 500g for a coarse duplicate. • Pulverizing to >95% <75μm and prepare a 25g bag of the pulp. • Pulp duplicate preparation. Two pulp bags with a minimum of 25 g each were prepared for the pulp duplicates. • Delivery of the pulps to the XRF laboratory. • Delivery of the coarse and the pulp rejects to WNP. • Control samples included: field, coarse and pulp duplicates, blanks, and 3 standards (low, medium and high grade). A sample of each control type was inserted at a rate of 1/100.

“In the second half of the year, after use of the low-temperature drying oven had been authorised, usually less than one week was needed to prepare the samples, compared to two weeks during the summer.

“In 2006, drying in the oven will be allowed since the temperature in the pulverizer is higher than in the oven.”

13.1.2 XRF Analyses in Ulaanbaatar

“The purpose of the XRF laboratory installed in Ulaanbaatar was to provide quick and reliable results used to: • give faster information to help direct drilling, and • check the assay results coming from the external laboratory as another means of QA/QC.

“One other long term objective was to train local employees to run the XRF laboratory and to validate its results in such a manner that in future they might be used as confidently as the results from external laboratories. 2005 was an experimental year, and the results to date indicate that this objective is far from being met.

“The chemist at the XRF laboratory has the following duties: • receive the listing of samples in the shipment by mail, • prepare the document for the export of the pulps, • send these documents to the Ministry, • get the signed documents back from the Ministry, • receive the pulp bags from Alex Stewart, • insert standards and blanks • rename the pulp duplicates according to the listing, • insert prepared standards with the shipment, • ship the pulps to Canada with an international air freight transporter (TNT or DHL), • prepare the XRF cells, • complete the XRF determinations, • send the daily results by mail with a file name ending in ‘P’ for preliminary, • insert the same standards in the cells as in the sample shipment to Actlab, • insert specific XRF standards with each batch analysed, • check the daily results and send the final results in a file specifying that the results are final.”

13.1.3 Activation Laboratories, Canada

“Eric Hoffman, PhD, P.Geo., General Manager of Activation Laboratories Ltd wrote this section of the report.

“Uranium was analyzed by delayed neutron counting (DNC) using a computer automated DNC system at the McMaster Nuclear Reactor. The system was custom designed and built by him (Eric Hoffman). High values were repeated by a lithium metaborate/tetraborate fusion to make a glass disc and then XRF using a Panalytical PW-1540 XRF.

“Vanadium and Barium were determined by pressed pellet XRF also using a Panalytical PW- 1540 XRF.

“Molybdenum was determined by aqua regia digestion and ICP using a Perkin Elmer Optima 3000 ICP.

“Fluor[ine] was done by lithium metaborate/tetraborate fusion and Specific Ion Electrode.”

13.2 QUALITY ASSURANCE AND QUALITY CONTROL PROGRAMS

13.2.1 Duplicates

13.2.1.1 Field Duplicates

“The field duplicate aims at monitoring the preparation in the field and the geological heterogeneity of a sample. Sample tags with numbers ending in 75 and 95 were used for field duplicate samples. A regular sample with a number ending in 75 was named FD1, and its field duplicate with sample number ending in 95 was named FD2, to facilitate understanding of the procedure.

“In the early shipments, some of the field duplicates were not prepared by the geologists due to a lack of understanding of the procedure. Field duplicates were spilt with a manual splitter as other samples in the batch. The initial sample was later split into two samples on a piece by piece basis. This procedure was set up as splitting – i.e. it is not done with an electric saw.

“With time, the procedure was changed and the initial sample was broken into very small pieces. Then all pieces in the bag were at once separated into two sample bags. But, as this procedure was not as effectively done as a real split, there was a bigger risk of heterogeneity, particularly for samples over a length of 0.5 m.”

13.2.1.2 Coarse Duplicates

“The coarse duplicate aims at monitoring the preparation at the laboratory as various assays are made from different pulps issued from the same coarse split. A selected sample sent to the Alex Stewart preparation laboratory is split after it has been dried and crushed. Two samples are prepared from the initial sample. For consistency, they are named CD1 for the regular sample and CD2 for the coarse duplicate.

“CD1 sample carries the initial number always ending in 15 to allow for a simple routine. The CD1 sample follows the same grinding, pulverizing and assaying process as all the other samples in the batch.

“CD2 is prepared in the Alex Stewart preparation laboratory in Ulaanbaatar. The crushed initial sample is split and two pulps are prepared separately out of these two coarse fractions. A CD2 sample uses the following number in the sequence ending in 25. Tags with numbers ending in 25 are marked as CD in the sample books and sent to the Alex Stewart preparation laboratory with the shipment listing and instruction letter. As the sample is not collected, the designation CD2 was introduced to allow the geologists and the technicians a quick check of the presence of the regular sample and its coarse duplicate tag in the same shipment. Confusion with this QA/QC procedure resulted in a risk of using the 25 tag for another regular sample.

“Alex Stewart preparation laboratory re-numbers the second fraction of the CD1 sample with the number ending in 25. This CD2 sample follows the same pulverizing and assaying process as all the other samples of the batch.”

13.2.1.3 Pulp Duplicates

“The pulp duplicate aims at monitoring the analytical precision as various assays are made from the same pulp. Two pulps are issued from the initial sample. The regular sample was named PD1 and the pulp duplicate PD2.

“The PD1 sample carries the initial number always ending in 45 for a simple routine. The PD1 sample follows the same assaying process as all other samples in the batch. The PD2 sample is prepared at the Alex Stewart preparation laboratory in Ulaanbaatar. The request form requires preparing two pulps of the PD1 regular sample and sending them with the rest of the pulps for shipment to our XRF laboratory. PD2 sample tags ending in 65 are sent with a complete listing in a special envelop to WNP’s technician in the XRF laboratory in Ulaanbaatar. The change of number for the second pulp of PD1 is made by this technician. The PD2 sample follows the same assaying process as all other samples in the batch.

“PD2 samples are noted as pulp duplicates in the sample book. The sample is not collected in the field. As the sample is not collected, the name of PD2 was introduced to allow for the geologists and the technicians a quick check of the presence of a regular sample and its pulp duplicate tag in the same shipment.”

13.2.2 Blanks and Standards - Selection

“For the Saddle Hills 2005 drill-program, blanks and standards were prepared using material from the Gurvanbulag deposit which had already been assayed several times and for which large quantities were available. They required validation.

“Fifteen standards were prepared at the Ulaanbaatar preparation laboratory but only 6 of them were immediately validated for the 2005 program. The 15 samples were chosen from among a set of 20kg samples collected by Gerald Harper in the Gurvanbulag area dumps to include 3 blank, 3 very-low grade, 3 low grade, 3 medium grade, 2 high grade and 1 very-high grade samples. Table 15 [Table 13.1] shows the samples selected from his database.

Table 13.1 Selection of Blank and Standard Samples

Initial Sample %U3O8 Duplicate Sample %U3O8 Variation Standard Type A07262-1 0.386 A07262-2 (1) & (2) 0.351 -8.87% Very high grade standard A07284-1 0.114 A07284-2 0.100 -12.80% High grade standard A07261-1 0.085 A07261-2 0.087 3.06% High grade standard A07255-1 0.028 A07255-2 0.029 5.04% Medium grade standard A07254-1 0.021 A07254-2 0.023 10.67% Medium grade standard A07291-1 0.019 A07291-2 0.019 3.18% Medium grade standard A07268-1 0.015 A07268-2 0.015 3.15% Low grade standard A07265-1 0.015 A07265-2 0.014 -2.44% Low grade standard A07263-1 (1) & (2) 0.014 A07263-2 0.013 -6.09% Low grade standard A07293-1 0.009 A07293-2 (1) & (2) 0.010 10.45% Very low grade standard A07273-1 0.009 A07273-2 0.009 0.13% Very low grade standard A07278-1 0.008 A07278-2 0.008 1.74% Very low grade standard A07426-1 0.001 Blank A07445-1 0.001 Blank A07446-1 0.001 Blank

“Selected samples were pulverized, homogenized and separated into separate 25g samples at the Ulaanbaatar Alex Stewart preparation laboratory.

“One pulp bag of each type was immediately sent for assay to Actlab and a second sample to Becquerel, both in Canada. Results of the new assays from Actlab and Becquerel were compared to the initial assays and checks (Table 16) [Table 13.2].

Table 13.2 Results of New Assays on Control Samples

Standard Initial Sample Initial GH Act Lab Bequerel Number Number ppmU ppm U ppm U % U3O8 ppm U % U3O8 A06001 A07254-1 178 197 180 0.021 219 0.026 A06002 A07255-1 238 250 233 0.028 234 0.028 A06003 A07261-1 720 742 747 0.088 717 0.085 A06004 A07262-1 3270 2980 2990 0.353 2860 0.337 A06005 A07263-1(1)&(2) 115 108 116 0.014 112 0.013 A06006 A07265-1 123 120 120 0.014 116 0.014 A06007 A07268-1 127 131 130 0.015 126 0.015 A06008 A07278-1 69 70 69 0.008 65 0.008 A06009 A07273-1 74 74 69 0.008 64 0.008 A06010 A07284-1 969 845 850 0.100 820 0.097 A06011 A07291-1 157 162 163 0.019 157 0.019 A06012 A07293-1 79 88 82 0.010 77 0.009 A06013 A07426-1 5 5 0.001 5 0.001 A06014 A07445-1 5 5 0.001 5 0.001 A06015 A07446-1 5 5 0.001 5 0.001

“The new Actlab results’ correlation with initial assay results was r2=0.9989 with Gerald Harper’s 1st initial assay and r2=0.9999 with Gerald Harper’s 2nd initial assay and the Actlab result (initial assays were also at Actlab). Overall, the new check assay results agree with the original assays to within 10%.

“Six of the foregoing samples were selected for the round robin test A series of random numbers was prepared for 12 bags of each of the pre-selected standards. Six of them were sent to Actlab and 6 to Becquerel.”

13.2.2.1 First Round Robin – June 2005

“The first round robin of June 2005 consisted of sending 6 samples of each of the selected blank/standard to Actlab and another set of 6 of each to Becquerel. The 6 samples were assayed with a required check. Then the same pulps were to be shipped by the first laboratory to the other one for a second assay on the same pulp. This step was not completed. The new numbers used for the round robin were randomly chosen to avoid repeating the same suite of samples several times.

“Ideally, the round robin should include more laboratories and more assays. The few facilities which can assay for uranium limit the validation program to the above. The remaining samples of validated standards were stored for immediate use in the program. The other 9 samples initially tested were stored separately for future validation.

“A summary of the results of the first round robin for blank (BL), very low grade (VLG), low grade (LG), medium grade (MG), high grade (HG) and very high grade (VHG) standards are presented in Table 17 [Table 13.3].

“After a comparison between the assays to ensure all 30 assay results were meaningful, the average value was initially adopted as the standard / blank value and the +2 and -2 standard deviation limits represented the acceptable limits for assay results during the QA/QC program.”

Table 13.3 Statistics for Standard Samples During First Round Robin

Sample Type # Samples Average Standard -2 Standard +2 Standard Number (%U3O8) Deviation Deviations Deviations (%U3O8) (%U3O8) (%U3O8) A06014 BL 15 0.001 0.000 0.000 0.001 A06009 VGL 16 0.008 0.0003 0.007 0.009 A06006 LG 16 0.014 0.0002 0.013 0.014 A06002 MG 16 0.028 0.001 0.026 0.030 A06003 HG 16 0.086 0.002 0.082 0.089 A06004 VHG 16 0.358 0.014 0.330 0.387

13.2.2.2 Second Round Robin – October 2005

“On October 24, in view of discrepancies between XRF results and validated standard values, a second series of standard pulps was sent out for a second round robin, starting with Becquerel. The round robin incorporated the same standards plus a second high grade unvalidated standard that had been used by mistake in the XRF laboratory.

“The standards were renumbered and six pulps were sent for each standard except for the blank, with only three. Becquerel analysed each pulp three times. Becquerel then sent the pulps to Actlab.

“A summary of the results of the second round robin for blank (BL), very low grade (VLG), low grade (LG), medium grade (MG), high grades (HG and HG2) and very high grade (VHG) standards are presented in Table 18 [Table 13.4].

Table 13.4 Statistics for Standard Samples During Second Round Robin

Sample Type # Samples Average Standard -2 Standard +2 Standard Number (%U3O8) Deviation Deviations Deviations (%U3O8) (%U3O8) (%U3O8) A06014 BL 18 0.001 0.000 0.001 0.001 A06009 VGL 36 0.008 0.0002 0.008 0.009 A06006 LG 36 0.014 0.0003 0.014 0.015 A06002 MG 36 0.028 0.001 0.026 0.030 A06003 HG 36 0.088 0.002 0.085 0.092 A06010 HG2 35 0.100 0.004 0.096 0.103 A06004 VHG 36 0.368 0.020 0.328 0.409

“The comparison between the round robin #1 and #2 results show a variation of 1 to 3% in the average grades for each standard. Both round robin results were combined to give the accepted values (averages) and the bracket (standard deviations) for each standard (Table 19) [Table 13.5].”

Table 13.5 Combined Statistics for the Standard Samples

Sample Type # Samples Average Standard -2 Standard +2 Standard Number (%U3O8) Deviation Deviations Deviations (%U3O8) (%U3O8) (%U3O8) A06014 BL 33 0.001 0.000 0.000 0.001 A06009 VGL 52 0.008 0.0003 0.008 0.009 A06006 LG 52 0.014 0.0003 0.014 0.015 A06002 MG 52 0.028 0.001 0.026 0.030 A06003 HG 52 0.088 0.002 0.083 0.092 A06010 HG2 37 0.100 0.002 0.096 0.103 A06004 VHG 52 0.365 0.019 0.327 0.403

13.2.3 Blanks and Standards - Insertion

“Insertion of standards and blanks occurs in the XRF laboratory in Ulaanbaatar. Pulp bags were first prepared, and the tags sent to the WNP XRF facility together with the shipment papers.

“To allow for the introduction of blank and standard samples in the series of pulps sent to Actlab in Canada, a simple procedure was followed: • Sample tags with numbers ending in 05 were noted as blank samples. Note that the blank was inserted at the pulp stage. There was no coarse geological blank inserted in the field during the 2005 program. This has been corrected for the 2006 program and re-sampling of some mineralized zones of the 2005 program with the geological blank is in process. • Sample tags with numbers ending in 35 were alternately used for very high or high grade standards. • Sample tags with numbers ending in 55 were noted as medium grade standards. • Sample tags with numbers ending in 85 were alternately used for very-low or low grade standards.

“The XRF technician was given a special listing for insertion to ensure that all QA/QC samples, and particularly blanks and standards, were introduced into the shipments to Canada. Several shipments to Actlab left the XRF laboratory without standards and blanks. The choice of standard was made by the logging geologist, but this choice was not always respected by the XRF laboratory technician. Furthermore, the standard inserted in the shipment to Actlab was often different from the one in the batch measured in Ulaanbaatar with the XRF analyzer.”

13.2.4 Re-assaying Procedures

13.2.4.1 XRF results

“For each shipment batch, the XRF values are received first, and they give a good idea of the future assay results of the samples. XRF results are usually available within 10 days after the samples leave the camp. The delay in sample preparation at Alex Stewart decreased with the use of a low temperature drying oven.

“Immediately after receiving the daily XRF results a rapid check is made on:

• sample numbers versus shipment list, • standards and blanks results, • field duplicates, • coarse duplicates, • pulp duplicates,

“Other controls, especially in mineralized zones, include comparing the quick logs, geological logs and radiometric logs with XRF results. When something does not appear normal, reassaying is immediately requested from the XRF technician. As a first step, the same cells are re-assayed to ensure that the suspected problem is not due to labelling or inversion mistakes. The required reassaying batch includes the suspicious sample and others.

“If the results the XRF re-assays are equivalent to the initial results, there are two options, depending on the XRF laboratory work load: • wait for the Actlab assays and compare them with the XRF results, • prepare a new XRF cell with the pulp rejects.

“In the first case, if the XRF and Actlab results are comparable and there is no other suspicion, the problem is deemed solved and does not require any follow-up. If the results of XRF and Actlab are comparable, but there is still doubt, a complete repreparation from the sample coarse reject is requested from Alex Stewart. This happened in shipment SH-2005-08 with a suspicion of inversion between two coarse duplicates in the shipment. The re-preparation confirmed the inversion made at the sample preparation level.

“When Actlab’s results and the initial XRF results differ for the samples, the first step is to re- assay these samples and a few additional ones in the same batch in the XRF laboratory in their initial cells. Most differences are solved at this stage.

“A second step is to prepare new cells from the original pulp and from a few additional ones in the same batch in case of a discrepancy caused by wrong labelling of the cells. In 2005, many analytical errors were due to inversions or wrong labelling of samples in WNP’s XRF laboratory. One function of the XRF laboratory is to check the quality of the external laboratories, but in reality it was often the other way around.”

13.2.4.2 ACTLAB results

“Upon receiving Actlab’s preliminary result, after about one month, they are compared with the XRF results. Duplicates are checked and the listing is compared with the shipment list. Five types of errors were common:

• For earlier shipments, sample numbers in assay reports were often different from WNP’s listing. For instance, A09550 might be written 9550 or A9550 or A-9550. These differences continued until it was realized that the listing was not sent to Actlab by the XRF laboratory technician as required in the procedure. Consequently, Actlab had to enter all sample numbers by hand and this caused mistakes. To avoid this problem, the person responsible for the shipment from camp now sends an advance copy of the request letter with the sample listing by mail directly to Actlab. These documents are also sent by courier with the samples, but this early mail transmission provides a back-up in case their insertion in the package is forgotten.

• Too few results. Either the missing samples were not part of the shipment or the results were missing. Two types of failures were identified. Either the XRF technician did not insert standard or pulp duplicates with the samples shipped, or Actlab reported a wrong number. Checks were required each time from Actlab and the error was quickly found and a new report issued. • Too many results. Samples in the report were not part of the shipment. Again, reporting was the most common problem. In each case, Actlab was asked to check and issue a new report. • A reporting error between two pages due to a gap in the Excel file between sample numbers and results. This happened twice in 2005. Each time, Actlab was asked to revise the report and issue a correct one. • A difference was often noted in the decimals between preliminary and final results, with generally two decimal numbers given in the preliminary results and only one in the final. This was discovered during an internal audit of the database. No satisfactory explanation was given by Actlab, but the laboratory is still working on the issue.

“Most of the time, the differences between the values reported by the two laboratories resulted from poorly prepared cells and assay mistakes in WNP’s XRF laboratory, due to human error and not to the XRF Niton spectrometer. To improve the quality of WNP’s XRF laboratory, a new chemist was hired in 2006 to replace the first one.”

13.2.4.3 Special Re-assaying Procedure

“A special re-assaying procedure is in effect at WNP: shipments SH-2005-13 and SH- 2005-19 were new pulp assays requested by Gerald Harper, WNP’s Vice President of Exploration, without any intervention of the staff involved in the routine of normal shipments. The results of the checks in these special shipments do not differ from the original values.”

13.2.5 Re-Sampling Procedure

“The barren core and the remaining half the mineralised intersections are stored in the original core boxes at the camp site. Core boxes are piled hole by hole. Should resampling be necessary, approval of the Project Manager or the Vice President Exploration is required. Once core is stored, nobody is authorized to go to the storage area and handle stored core without good reason. Core re-sampling would be limited to one quarter of the core to keep at least that much again of the original drilled intersection. In 2005, re-sampling was not once requested, and at least half of all sampled core remains available in the storage area.”

13.2.6 Record Keeping for Traceability

13.2.6.1 Drill Logs

“Geological, geotechnical, SG and radiometric drill logs are completed by hand writing on paper in the core shed and the information is reported on an Excel spreadsheet. The initial log form includes sample numbers but assay results are only reported later on the log after their validation.

“The original paper drill log is filed in a binder, starting with drillhole number GCE5001. The original may be consulted at all times and is never to be removed. If taking the log away in the field became necessary, a copy was made. Corrections to the original log were sometimes necessary. They were made either on a new log form or by correcting a photocopy of the

original log form and filing them together. The original hand-written paper log is never discarded. The original hand-written log is also copied and sent to Ulaan Baatar (UB) for safety purpose.

“To validate the logs, there are three steps:

• Initial entry of the log data in a spreadsheet: the spreadsheet’s file name bears the number and the type of the log followed by the initial of the person responsible for the data entry (as in “GCE5067 Geotechnical log B2D.xls”). The information on the name of person who entered the data and the date is noted at the bottom of the page for the geotechnical, SG and radiometric logs, in the header of the geological logs. • A first check is made by the geologist in charge of the drillhole. At this point, the name of the file has changed to replace the initials of the data entry clerk by the letters ‘chk’ and the initial of the geologist (as in “GCE5067 Geotechnical log Check ME.xls”). The geologist corrected the different mistakes in the data entry and coding errors. The different graphs (recovery % and radiometric measures) are also checked. The information on the name of person who checked the data and the date is noted at the bottom of the page for the geotechnical, SG and radiometric logs, in the header of the geological logs. • Validation of the logs was done by the project manager of the QA/QC program. Although the objective remains that the logging geologist be responsible for his/her drillhole from start of drilling to validation of the drillhole, this was not achieved in 2005. The corresponding file is finished by the letters val (as in “GCE5067 Geotechnical log val.xls”). The information on the name of the person who validated the data and the date is noted at the bottom of the page for the geotechnical, SG and radiometric logs, in the header of the geological logs. It is only after validation that the geological or geotechnical information contained in the log is entered in the database.

“The corresponding prints of the various Excel spreadsheets also get filed in the binder. When assays are added, the new print of the Excel spreadsheet is added in the binder. The computer file is copied to the central filing system at camp and sent to management by e-mail. Back-ups are regularly sent to Ulaanbaatar.”

13.2.6.2 XRF analysis results

“Copies of the XRF analysis results are received from WNP’s XRF laboratory in Ulaanbaatar via e-mail. The results are plotted on a sample result form and entered in the database. The XRF results are filed with the other logs and in a specific binder for analytical results on a shipment by shipment basis.”

13.2.6.3 Assays

“All paper work related to the sample shipments are filed in the assay binder. Each shipment is identified and separated from the previous and following ones by dividers.

“Result reports from the Actlab laboratory in Canada are received from the laboratory by e- mails also sent to the Vancouver office and to Vice President Exploration in Toronto. Preliminary results are received within 8 days to 65 days from shipment out of site. Final non certified results may take as long as 100 days. The signed assay certificates are then sent by

these laboratories to the Vice President Exploration office in Toronto. Paper copies are finally sent to the field office where they are filed in the result binder.”

13.2.6.4 Drillers reports

“Drillers’ reports are checked by the chief geologist and/or the project manager. Driller’s report contain details on all activities at the drill (hours worked, length drilled, core size etc… including incidents). They are used to check the drill company invoices but also to back-up the evaluation of incidents on core integrity and therefore on ultimate quality of the data. They are filed in specific binder at the exploration office.”

13.2.7 Data Storage and Security

“Security procedures were established to ensure that in any circumstances, no or only the minimum data would get lost. Duplication of paper and computer filing as well as duplication of storage facilities participate to protecting the integrity of data.

13.2.7.1 Paper data

“Paper data is filed in the Saddle Hills office in ring binders as described above. Completed logs are regularly sent with trucks returning from Saddle Hills to Ulaanbaatar. They are stored with the other technical files in WNP’s UB office. At this point, no other paper copy was deemed necessary. Other offices receive copies of the electronic files.”

13.2.7.2 Computer data

“All paper work was immediately entered into Excel spreadsheets. Data from these spreadsheets were used on site to update the Excel database through simple data extraction. The Excel database was in turn used to produce the Gemcom database. These files are securely stored and provided to other involved consultants.

“All new computer information is regularly backed-up on various computers and hard drives. A full project back-up is made at least once a month on DVDs or CDs. One copy is at the WNP’s Ulaanbaatar office for storage together with the paper copies of logs. Another copy is provided to management for storage in Canada. The update of the database is also regularly sent to a consultant.”

13.3 COMPILATION OF RUSSIAN ANALYTICAL DATA

“There is limited documentation as to the sampling and QA/QC protocols used by the Russians. The available reports verify that their analytical and estimation methods followed those approved by the State Committee on Mineral Reserves (GKZ) of the USSR in 1982. At Gurvanbulag, the grade and thickness of the mineralized intervals were determined primarily from gamma radiation logs (Kiselev et al., 1985). The Russians devoted considerable attention to the evaluation of gamma radiation to be used as a proxy for direct assay measurements of the uranium grades and found that their results were acceptably consistent for this purpose (Rogov and Urchenko, 1987). No records exist of original assays or data used in these comparisons, and the data have been incorporated as they appear on detailed cross-sections prepared by the Russians and digitized by WNP.”

14.0 DATA VERIFICATION

Data verification is described in SRK (2006), as follows:

14.1 CONTROL OF ASSAY RESULTS

“WNP maintained very thorough quality assurance protocols throughout their analytical program, monitoring the Alex Stewart prep lab and their own XRF lab in Ulaanbaatar, as well as, their primary lab in Canada, ActLabs, as described in Section 12.2 [see Section 13.2]. This process identified numerous errors from sample mix-ups, to a lack of insertion of control samples by the prep lab, and low battery charge in the XRF in Ulaanbaatar.

“WNP’s QA/QC program was audited on site by Dr. Barry Smee, PGeo, and the author in February, 2006. Key findings of Dr. Smee’s audit are summarized here and his report is included in full as Appendix A, at the end of this report. [See SRK (2006)]. Regarding the use of control samples at Saddle Hills, Dr. Smee made four important observations:

• Blanks were submitted as pulps to the analytical lab and therefore missed the sample prep stage. As the sample prep stage is the most vulnerable to sample contamination, these blanks do not serve their intended purpose. • Due to mistakes in the Alex Stewart Laboratory that were not caught in time, only one standard was inserted for every 100 samples. This resulted in approximately 40% of the drillholes not having any standards to confirm the accuracy of the results. • The standards prepared by WNP are not supported by an adequate ‘round robin’ analysis to properly characterize their mean values and standard deviations. • The selected standards are not at appropriate grades to support the principal economic thresholds of the deposit.

“Since Dr. Smee’s visit, WNP have made every effort to refine their quality control practices. Most importantly, they have implemented a comprehensive re-assay program to confirm the quality of the analyses lacking appropriate control data (blanks and standards). These re-assays are being conducted using 4 CANMET standards at more appropriate grades for the Gurvanbulag deposit.

“This author [Christopher Lee] has not reviewed the results of this re-assay program.”

14.2 ASSAY DATABASE

“The original assay database, maintained on-site, was kept in Excel spreadsheets (see WNP 2005 annual report for a detailed description). These spreadsheets contained a comprehensive and complex record of all information regarding sample location, results (original and repeat analyses), shipment details, quality control data, comments regarding sample mix-ups and other errors, as well as, corrective measures. Maintenance of such a complicated spreadsheet was a full-time job, and was only attended to (and understood by) a single person. Multiple edits and updates to the master file made this system prone to error, and when errors were found, they were difficult to trace.

“Prior to the transfer of the assay data to SRK, a new relational SQL database was constructed from scratch by Maxwell Geoservices in Vancouver. Maxwell’s database was constructed from

original assay certificates, as received from Actlabs, and all QA/QC data were reviewed by Caroline Gilson, PGeo (QP), at Maxwell. The results of Maxwell’s QAQC review are summarized below, and their report is appended as Appendix B, at the end of this report [see SRK (2006)]:

“Standards and blanks: • U3O8 and U analyses display good accuracy • F analyses are variable, possibly to due poor instrument calibration for higher concentrations • Mo analyses show reasonable accuracy • V analyses show variable but acceptable accuracy.

“Duplicates and Repeats • All analyses exhibit strong repeatability for all analytes

“Ms. Gilson maintained regular contact with Doug Blanchflower, PGeo, and the author of this report (C.Lee), who carefully reviewed and validated the database construction. SRK is confident that the assay database is in good order, and the data of sufficient quality for resource estimation.”

15.0 ADJACENT PROPERTIES

SRK (2006) concluded that there are no adjacent properties relevant to the contents of its report.

Micon concurs with this conclusion for the purpose of its preliminary economic assessment.

16.0 MINERAL PROCESSING AND METALLUGICAL TESTING

Western Prospector/Emeelt Mines retained Melis Engineering Ltd. (Melis) to direct a program of metallurgical testwork on samples from Gurvanbulag, to develop a preliminary processing flowsheet and, specifically for the purpose of this preliminary economic assessment, to provide estimates of capital and operating costs for a processing plant. Western Prospector/Emeelt Mines provided Melis with copies of the translated reports of Russian metallurgists describing the results of their testwork on the deposit.

In connection with the work by Melis, Bruce Fielder, P.Eng., Principal Process Engineer, visited the Gurvanbulag site in October 13-25, 2006 and June 15-20, 2007.

16.1 METALLURGICAL TESTWORK

Metallurgical testwork for the Gurvanbulag project was carried out under the direction of Melis at SGS Lakefield Research Limited (SGS Lakefield) on samples of drill core received at the laboratory in June, 2006. The samples represent the mineralization as obtained during drilling up to May, 2006. It should be noted that access to the underground exposure, as a result of mine dewatering, will allow collection of samples that have not subjected to differential washing out during drilling.

The testwork program included grindability, leaching, liquid/solid separation, solvent extraction, precipitation and tailings preparation. It was completed in December, 2006 and was described in the Melis report, Summary of Metallurgy, dated February 14, 2007 (Melis (2007a).

The summary of the Melis (2007a) report is provided below. For reference to the locations of mineralized material from specific resource blocks within the Gurvanbulag Central Zone, as referred to in the Russian resource classification within the deposit, see SRK (2006) and Harper (2005).

16.1.1 Composite Preparation

“Twenty-three drill core samples of the Gurvanbulag mineralization were received at Lakefield in mid-June 2006. The samples were referenced and separated as Block 1-6, the shallower block with apparently a greater amount of coffinite, and Blocks 2-1 and 2-5, a deeper area with less coffinite.

“Five test composites were prepared from these samples. The average grade composites assayed 0.26% U3O8 to 0.33% U3O8, a high grade composite assays 4.74% U3O8 and a low grade composite prepared for a bottle roll heap leach/underground stope leach test assayed 0.058% U3O8.

“The following composites were prepared for testing:

“Composite 1AG – blend of samples from Block 1-6 to provide a composite with an average grade of 0.26% U3O8,

“Composite 1M – blend of four selected samples from Block 1-6 to provide a high molybdenum composite containing 0.35% U3O8 and 0.012% Mo,

“Composite 2AG – blend of samples from Blocks 2-1 and 2-5 to provide a composite with an average grade of 0.33% U3O8,

“Composite 2HG – blend of two high grade samples from Blocks 2-1 and 2-5 to provide a high grade composite with a grade of 4.47% U3O8, and

“Composite 2LG – blend of selected low grade samples from Blocks 2-1 and 2-5 to provide a low grade composite with a grade of 0.058% U3O8.

“Significant assays for these composites are summarized in Table 1 [Table 16.1, below].”

Table 16.1 Test Composites – Detailed Analyses (%)

Analyte IAG-VAR IM-VAR 2AG-VAR 2HG-VAR 2LG-VAR U3O8 0.26 0.36 0.33 4.74 0.058 As 0.012 0.009 0.007 0.038 0.008 Mo 0.0058 0.012 0.0011 <0.001 <0.001 Ni 0.004 <0.002 <0.002 <0.002 <0.002 Se <0.003 <0.003 <0.003 <0.003 <0.003

“The uranium isotope in Composite 1AG was 0.72% U235/U (w/w), exactly the value expected in nature. This confirms that the deposit value calculations assuming a U235 proportion of 0.72% are valid and that the uranium deposit is naturally occurring.

“The Bond Work Index or value for the Gurvanbulag mineralization was measured at 23.5 kWh/t on the average grade composite indicating that it is relatively hard mineralization. This high BWI conflicts with the expectations of project geologists who have said that the uranium occurs primarily in the softer rock of the uranium zone. Samples collected for the Feasibility Study [i.e., the study planned to be undertaken subsequent to the Micon preliminary economic assessment] are expected to be more representative of the mineralization lithology due to better definition of the resource during the ongoing drilling program. In that phase, variability testing will specify the hardness of the mineralization in each of the rock types for inclusion in the block model.

“Acid-base accounting and net acid generation tests were performed on Composite 1AG. Acid- base accounting tests indicated a CO3 NP [neutralization potential] to AP [acid generation potential] ratio of 1.1, indicating an uncertain potential for acid generation. Net acid generation potential tests indicated an acid production equivalent of 0.30 kg H2SO4/t of Composite 1AG. Because of the low net acid generation potential, no detectable sulphide concentration, and the low calculated CO3 net acid generation, it is expected that Composite 1AG would generate a near neutral pH (6.2) environment with little buffering capability.”

16.1.2 Leach Tests

“Nine optimization and one bottle roll test were performed on Composite 1AG. After optimum leach conditions were selected, five variability leach test were performed, one on each of the five composites prepared. Leach test conditions and results are summarized in Table 2 [Table 16.2, below].”

Table 16.2 Summary of Leach Test Conditions and Results

Test No. Test Conditions Reagent Additions Test Results 3+ Temp. Target FA Avg. ORP Grind H2SO4 NaClO3 Fe Calculated 24-h U3O8 24-h Leach o ( C) (g H2SO4/L) (mV, K80 (kg/t) (kg/t) (kg/t) Head extraction Residue Ag/AgCl) (μ) (% U3O8) (%) (% U3O8) 1AG-S1 50 15 657 134 40.0 12.0 - 0.233 96.9 0.0073 1AG-S2 50 15 507 134 33.8 - 5.0 0.226 96.3 0.0083 1AG-OPT1 50 5 545 134 13.6 1.0 - 0.211 96.0 0.0084 1AG-OPT2 50 15 534 134 36.9 - - 0.204 97.1 0.0060 1AG-OPT3 20 10 510 134 28.9 6.8 - 0.211 95.3 0.0099 1AG-OPT4 35 10 560 134 29.0 3.1 - 0.223 96.0 0.0088 1AG-OPT5 50 10 533 134 27.8 0.7 2.0 0.235 97.0 0.0071 1AG-OPT6 50 10 509 178 29.2 - - 0.231 96.4 0.0084 1AG-OPT7 50 10 522 227 26.9 - - 0.212 95.2 0.0101 BRL-1 23.7 9.31 5861 ~1,600 48.1 - - 0.061 88.72 0.0068 1AG-VAR 50 10 503 134 39.0 0.7 - 0.239 95.5 0.0107 1M-VAR 50 10 521 ~120 36.7 - - 0.314 94.4 0.0177 2AG-VAR 50 10 556 140 30.7 - - 0.320 97.2 0.0088 2HG-VAR 50 10 596 121 142.4 - - 4.715 95.0 0.2358 2LG-VAR 50 10 615 ~90 32.3 - - 0.041 82.8 0.0071 1 Final (16 day) free acid [FA] and [oxidation reduction potential] ORP. 2 Final (16 day) U3O8 extraction.

“Nine scoping and optimization tests, identified as 1AG-OPT1 through 1AG-OPT7, were performed on Composite 1AG, a blended composite grading 0.26% U3O8. All the uranium extractions obtained in the scoping and optimization tests were higher than 95% and all but two higher than 96%. The two tests with the highest 24 hour uranium extractions were Test Nos. 1AG-OPT2 and 1AG-OPT5, with 24 hour uranium extractions of 97.1% and 97.0%, respectively.

“Test No. 1AG-OPT2, the test with the highest uranium extraction, required no oxidant addition at all. Test No. 1AG-OPT5, with the second highest 24 hour uranium extraction, required only 3+ 0.7 kg NaClO3/t and 2 g Fe /t.

“The scoping and optimization leach tests suggest that to obtain a 24 hour leach extraction of approximately 97%:

• The primary grind should have a K80 of approximately 134 μm,

• The leach temperature should be approximately 50oC,

• The free acid concentration should be a minimum of 10 g H2SO4/L, and

• The oxidation/reduction potential should be in the range of 500 mV-535 mV.

“Achieving an average oxidation/reduction potential of 500 mV-535 mV can require as little 3+ oxidant as 0.7 g NaClO3/t and 2.0 kg Fe /t.

“When leaching Composite 1AG, averaged over every scoping and optimization test, a U3O8 extraction of 95% can be obtained after 7 to 8 hours of leaching.

“One bottle roll leach test, identified as BRL-1, performed on Variability Composite 2LG-VAR with a calculated head of 0.061% U3O8, achieved a uranium extraction of 88.7% after 16 days. Though unconfirmed, it is possible extraction was still occurring at the end of the 16 day test. Extrapolation from the test results suggests that an extraction of 95% may have been achieved sometime after 35 days of testing. If confirmed, heap leaching or underground stope leaching may thus prove to be an economic option for low grade ore.

“Five variability leach tests, identified by the suffix “-VAR”, were performed on variability composites 1AG, 1M, 2AG, 2HG and 2LG. Variability test conditions were based on those o used in Test No. 1AG-OPT5: 50 C temperature, 10 g H2SO4/L free acid, 500 mV ORP, and a grind K80 of roughly 134 μm. No ferric sulphate was added in the variability tests.

“With the exception of Test Nos. 1M-VAR and 2LG-VAR, extractions of 95.0% or higher were achieved in all the variability composite leach tests. The extraction for Variability Composite 2LG-VAR started and remained much lower than those for the remaining variability composites, likely due to the quite low calculated head of this composite, 0.041% U3O8. It should also be noted that there appeared to be some difficulty with this test, as indicated by the negative weight loss of -0.2% during leaching. Test No. 1M-VAR had a relatively low extraction of 94.4%; this was likely due to the 0.012% Mo in this composite competing with uranium for available oxidant.”

16.1.3 Settling and Flocculation Tests

“A series of settling and flocculation tests were performed on neutral (pH 7.9) feed slurry and an acidic leach residue slurry. The purpose of the tests was to determine the optimum flocculant and thickener unit area for neutral feed slurry and to determine the thickener unit area for acidic slurry.

“Magnafloc 155 was selected for flocculating the neutral pH slurry as it gave the largest and fastest settling flocs. The addition of coagulant to the supernatant was required to produce a clear supernatant. Settling tests were conducted at Magnafloc 155 dosages of 30 g/t, 40 g/t and 51 g/t. There was a significant decrease in thickener underflow unit area between dosages of 40 g/t and 51 g/t, (0.11 m2/t/d to 0.05 m2/t/d) with no significant decrease in the underflow solids density (57.3% solids (w/w) at 40 g/t and 57.2% solids (w/w) at 51 g/t).

“Based upon operating experience at Canadian uranium mills, Magnafloc 351 was selected as the optimal flocculant for use on acidic leach residue, i.e., counter current decantation (CCD) feed. Settling tests were conducted at Magnafloc 351 dosages of 55 g/t, 66 g/t, and 83 g/t. There was a significant decrease in thickener underflow unit area between dosages of 55 g/t and 66 g/t, (0.47 m2/t/d to 0.27 m2/t/d) with a slight increase in the underflow solids density (52.6% solids (w/w) at 55 g/t and 53.6% solids (w/w) at 66 g/t). Increasing the flocculant dosage to 83 g/t decreased the thickener underflow unit area to 0.18 m2/t/d. In addition, clear supernatant was produced with Magnafloc 351 dosages of 55 g/t and 66 g/t, an important consideration for solvent extraction feed.”

16.1.4 Solvent Extraction Tests

“Extraction and stripping shake out tests were conducted on clarified pregnant leach solution generated in the scoping, optimization and variability leach tests. The purpose of the tests was to check extraction and stripping efficiencies and generate raffinate for the production of tailings. The organic tested in extraction was a mixture of 2.5% Alamine 336, 2.5% isodecanol, and 95% Isopar M. Stripping solutions tested were ammonium sulphate and strong sulphuric acid.

“The steep initial slope of the extraction isotherm indicated that extraction from the pregnant leach solution into the organic was fast. The extraction isotherm is preliminary and serves mainly to indicate that no chemical difficulties should be expected on the extraction side of solvent extraction.

“The final uranium concentration achieved in the ammonium sulphate strip solution was lower than the 932 mg U3O8/L in the pregnant leach solution or the 825 mg U3O8/L in the pregnant organic solution, though solvent extraction did decrease the arsenic concentration from 12 mg As/L to less than 3 mg As/L and the molybdenum concentration from 13 mg Mo/L to an average 1.8 [mg] Mo/L. The lack of concentration of uranium in the ammonium sulphate strip solution was likely due to poor control of pH in the ammonium sulphate stripping test. As the use of ammonium sulphate stripping and precipitation also results in elevated ammonia concentrations in mill effluent, the use of that stripping chemistry may be counter indicated.

“The final uranium concentration in the strong sulphuric acid strip solution was 4,900 mg U3O8/L, higher than the 932 mg U3O8/L in the pregnant leach solution and the 825 mg U3O8/L in the pregnant organic solution. As well, the arsenic concentration decreased from 12 mg As/L in the feed solution to <3 mg As/L in the strip solution and the molybdenum concentration from 13 mg Mo/L in the feed solution to an average 6.4 mg Mo/L in the strip solution.

“Strong sulphuric acid stripping appears to be successful in concentrating uranium and decreasing arsenic concentration in the strip solution. Molybdenum concentration was decreased by approximately 50%.”

16.1.5 Uranyl Peroxide Precipitation

“A bench test of uranium precipitation from the loaded strip solution, precipitated as uranyl peroxide (UO4.4H2O) using the hydrogen peroxide precipitation process, was conducted. The feed for this test was loaded strip solution from the strong acid strip test. After sampling, there remained 430 mL of loaded strip solution with a grade of 6,130 mg U3O8/L. After precipitation, 2.15 g uranyl peroxide was obtained.

“The uranyl peroxide precipitation reaction had an efficiency of 99.7% for uranium and resulted in the barren strip containing 4.7 mg U3O8/L. This is a typical concentration of uranium in a barren strip solution and suggests that precipitation ceased because the uranium concentration in the barren strip solution reached an effective minimum.

“The concentration of impurities in the precipitate are listed on a percent uranium basis in Table 3 [Table 16.3, below] and compared with Typical (Cameco) Refinery Specifications, where available.”

Table 16.3 Summary of Uranyl Peroxide Precipitate Quality on a Percent Uranium Basis

Analyte % Analyte/ Typical (Cameco) Analyte % Typical (Cameco) Refinery % U Refinery Specifications Analyte/ Specifications % U Surcharge Reject Surcharge Reject Limit Limit Limit Limit Al 0.17 - - Mo 0.002 0.1 0.3 As <0.01 0.05 0.15 Na 0.01 1 2 Ba <0.0003 - - Ni 0.02 - - Be 0.0002 - - P 0.04 0.2 0.5 Bi 0.03 - - Pb 0.01 - - Ca 0.18 3 4 Sb 0.01 - - Cd 0.001 - - Se 0.005 - - Co 0.01 - - Sn 0.03 - - Cr 0.003 - - Sr 0.005 - - Cu 0.07 - - Ti 0.03 0.05 0.1 Fe 0.42 1 2 Tl 0.005 - - Hg 0.00005 - - Th <0.0029 0.5 2 K 0.05 - - V 0.003 0.1 0.5 Li 0.002 - - Y 0.01 - - Mg 0.16 3 4 Zn 0.01 - - Mn 0.02 - -

“All measured impurities were lower than the typical refinery reject and surcharge limits.

“On a percent uranium basis, molybdenum concentration decreased from 0.23% in the loaded strip solution to 0.017% in the uranyl peroxide precipitate. This is a sufficiently low molybdenum concentration to avoid penalty at the refinery, which typically sets a surcharge limit at 0.1% Mo/% U. A comparison of refinery specifications for other impurities showed that uranium product from the Gurvanbulag mineralization, processed with a sulphuric acid leach, solvent extraction with strong sulphuric acid strip and hydrogen peroxide precipitation is of acceptable quality.

“There was insufficient loaded strip solution from ammonium sulphate strip available to test ammonia precipitation of uranium.”

16.1.6 Tailings and Environmental Data

“Tailings were prepared from leach residue from tests on Composite 1AG and raffinate produced in the solvent extraction tests. The procedure used to generate tailings also generated treated effluent, which in a mill would be released to the environment. The tests performed on the tailings were:

• Elemental and radionuclide analysis on solids and supernatant,

• A detailed particle size analysis using a Malvern Instrument, and

• Aging tests, including elemental and radionuclide analysis of the supernatant on days 0, 14 and 30.

“The treated effluent was subjected to elemental and radionuclide analysis.

“Particle size analysis with the Malvern Instrument indicated that the K80 of the neutralized tailings was 126 μm.

“Aging tests on the prepared tailings were used to judge the quality of the supernatant released as the tailings consolidate over the first 30 days. As anticipated, radium assays (22.0 Bq Ra226/L on Day 0, 19.0 Bq Ra226/L on Day 14 and 19.0 Bq Ra226/L on Day 30) were all significantly above the Canadian maximum monthly arithmetic mean concentration of 0.37 Bq/L. This confirms that tailings supernatant must be treated to remove radium before release. The molybdenum concentration in the treated effluent was 29.8 mg Mo/L on Day 0, and so treatment to reduce the concentration of molybdenum in the supernatant will likely be required.

“Table 4 [Table 16.4, below] compares analytes of interest in the treated effluent with the maximum monthly arithmetic mean concentration limits for those analytes specified in the (Government of Saskatchewan) Mineral Industry Environmental Protection regulations and the (Government of Canada) Metal Mining Effluent Regulations.”

Table 16.4 Treated Effluent Analysis and Monthly Arithmetic Mean Concentration Discharge Limits

Analyte Units Treated Effluent Assay Maximum Monthly Arithmetic Mean Concentration Discharge Limits pH 7.8 6.0-9.5 TSS mg/L 48 15 As mg/L 0.0011 0.05 Cu mg/L 0.0066 0.3 Mo mg/L 2.41 1 Ni mg/L 0.0188 0.5 Pb mg/L 0.00282 0.2 Se mg/L <0.003 1 U mg/L 0.0129 2.5 Zn mg/L 0.0136 0.5 Pb210 Bq/L 0.2 0.92 Ra226 Bq/L 0.03 0.37 Th230 Bq/L 0.30 1.85 1 Not currently specified.

“The radium concentration in the treated effluent, 0.03 Bq Ra226/L, is 8% of the maximum monthly arithmetic mean concentration of 0.37 Bq Ra226/L. This indicates that a two stage radium removal treatment [refer to Appendix G of Melis (2007)] reduced the radium assays found in the tailings aging tests (22.0 Bq Ra226/L on Day 0) by over 99.8%. The arsenic concentration in the treated effluent was 0.011 mg As/L, a reduction of 98% from the 0.0575 mg As/L measured on Day 0 in the tailings aging tests.

“The TSS, at 48 mg/L, is higher than the maximum monthly arithmetic mean concentration limit of 15 mg/L. Under operating conditions, the use of sand filters for final clarification would be required to meet TSS discharge standards. With the probable exception of molybdenum, and the possible exception of selenium, for which discharge limits have not been defined, the concentrations of all other analytes in the treated effluent were below the maximum monthly arithmetic mean concentration limits.”

17.0 MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES

17.1 MINERAL RESOURCE ESTIMATION

SRK was retained by Western Prospector to prepare an independent mineral resource estimate for the Central Zone of the Gurvanbulag uranium deposit that is compliant with the mineral resource and mineral reserve definitions of CIM. The SRK report, entitled Independent Technical Report and Resource Estimate for the Gurvanbulag Deposit, Saddle Hills Uranium Project, Mongolia, is dated November 17, 2006 and was filed on SEDAR in June, 2007 (SRK (2006)).

As described in detail in SRK (2006), the resource estimate, itself, was based on a total of 1,458 holes of which 1,356 were Russian and 102 were drilled by Western Prospector/Emeelt Mines:

“The estimate was conducted on a dataset that combines both Russian data and new data collected by WNP in their 2005-06 drillhole program. Drilling was on-going at the time the estimate was conducted, but only those data from the first 118 WNP drillholes (completed on March, 18, 2006) were included in that estimate.

“The principal components of SRK’s estimation work included: • “Construction of a 3D geological model in Gemcom to constrain the data selection and interpolation processes; • “Detailed comparison of the Russian and WNP datasets, to ensure their compatibility and lack of bias between datasets; • “Comparison of 5 different estimation procedures • “Final resource estimates for reporting • “Classification of the resource.”

17.1.1.1 Grade Measurements

A total of 5,407 grade measurements from the Russian drilling and from the drilling program of Western Prospector/Emeelt Mines in 2005/06 were used in a design of a mineralized horizon straddling the obsidian-rich Gurvanbulag horizon. The Russian data were collected from Russian maps showing average grade intersections of mineralized portions of the deposit and the data from 2005/06 are based on assays from 0.3-1.84-m samples. Figure 17.1, reproduced from SRK (2006), shows the locations of Russian and Western Prospector/Emeelt Mines drill holes, with the outline of the Russian 1-06-C1 resource block, in which most of the new drilling is located.

Mineralized intersections reported on the Russian maps can be as narrow as 0.2 m or as wide as 38.0 m, as opposed to the maximum sample length of 1.84 m from the 2005/06 drilling. The Russian drill holes represent a combination of closely spaced (15 m) underground drilling and wider spaced (30-50 m) surface drilling while the majority of the samples (60%) from the 2005/06 drilling are approximately 1.0 m long.

Figure 17.1 Locations of Drill Holes and Area 1-06 C1

17.1.1.2 Compositing of Mineralized Intercepts

The 2005/06 drilling was directed towards sampling and assaying not only the higher grade zones but also the lower grade areas around the mineralized zones.

A total of 2,836 intercepts within the two mineralized domains were used in compositing for resource estimation. Composites of 0.5-m were used for resource estimation. Within the mineralized domains, a total of 12,405 composite assays were produced from 102 holes drilled in 2005/06 and 1,356 Russian drill holes.

The average thickness of the mineralized horizons is 2.3 m in the footwall domain and 1.7 m in the hanging wall. In unsampled areas, SRK injected an arbitrarily low value of 0.001% U3O8. The average thickness of the barren intervening obsidian layer is 6.1 m. SRK assumed that a lack of samples in any drill hole intercept indicated an absence of visible mineralization, and these intervals were assigned an arbitrarily low value of 0.001% U3O8. This resulted in there being no missing assays for any intercept within the modeled zones.

17.1.1.3 Geological Model

SRK (2006) describes the geological model for Gurvanbulag as follows:

“Uranium mineralization in the Gurvanbulag deposit is controlled by the stratigraphic contacts on either side of an obsidian-bearing horizon sandwiched between an ashy felsic tuff unit in its footwall and a massive ‘felsite’ in its hanging wall. […] There is some indication that the mineralization is controlled, at least in part, by structures that follow these contacts; however, the available data indicate that the stratigraphic contacts described above remain the dominant control on both the location of the structures and the mineralization. As such, the geological model was constructed on the basis of logged lithologies in the drillhole database, as opposed to the various structural elements (e.g. faults, fracture frequencies, etc.).

“As seen in Figure 11 [not provided in this report], there is a persistent barren horizon between the hanging wall (HW) and footwall (FW) zones that generally coincides with the obsidian- bearing layer. This boundary is not absolute, however, and uranium mineralization does bleed into the obsidian-bearing layer, locally. The HW and FW mineralized domains were designed to honour these relationships wherever possible. In other words, the logged intercepts of the obsidian layer were used to define the upper or lower limits of the mineralized domains, unless the layer contained grade, in which case the domain boundaries were modified to capture this grade.

“The mineralized domain intervals for each drillhole were picked using the following criteria:

“• The mineralized domains are concentrated only on those mineralized intercepts occurring at, or near, the upper and lower contacts of the obsidian-bearing horizon - i.e. the model does not include any steeply dipping mineralized veins or structures, or mineralized horizons that are known to occur at deeper stratigraphic levels. “• Domain thicknesses were determined by either the width of the mineralized interval above 0.07% U3O8, or a true thickness of 1.5 metres, which ever is greatest. “• For drillholes with no grade data, or any missing sampled intervals, a grade of 0.001% U3O8 was assigned over a true thickness of 1.5 metres. “• Unless otherwise indicated by grade measurements, the obsidian layer contacts were used as either the upper or lower contacts of the domains.

“The imposed continuity, albeit at much lower grades, is considered reasonable since the stratigraphic contacts are considered to have been the primary pathway control on the mineralization. SRK is of the opinion that, given the nuggety nature of the mineralization, this approach is more practical and realistic than modelling numerous, discontinuous high grade lenses.

“The zones are undulating and generally gently south-easterly dipping (15° dip). The mineralized zones can be very close to each other (i.e. the obsidian layer is almost nonexistent) or they can be separated by 5-10m.

“The dip of the mineralized horizons is fairly uniform across the deposit, apart from a well- defined conical depression in the HW and FW surfaces, where the surfaces drop approximately 40 metres below their projected horizon […]. This depression coincides with the intersection

between major northeast and north trending faults mapped by the Russians; however, the cause of this peculiar geometry is unclear.

“The two zones are slightly more variable along strike (Figure 14) [not provided in this report] at a local scale, although this is partly an expected consequence of their shallow dip. In places, the strike can change fairly abruptly, along with dip, apparently reflecting minor offsets in the stratigraphy. The continuity of mineralization across these boundaries suggests that these minor offsets predate the mineralization.

“The HW and FW domains were used to constrain 0.5 metre composites of all mineralized intervals, which were then used in the resource estimation. More than 90% of the composites are derived from the Russian data which are average grade intercepts and do not reflect the true grade variability of the mineralization. The resulting 0.5 metre composites produced for this estimate exhibit a similar lack of grade variability compared to the in situ grades.”

SRK (2006) undertook statistical analysis of the grade data from the two sets of drill hole data.

17.1.2 Resource Estimation Methodology

SRK (2006) describes the methodologies utilized for resource estimation as follows:

“The HW and FW mineralized domains are generally gently dipping with slight undulations, but with locally abrupt changes in dip. These abrupt changes in dip present some difficulty in interpolation for block estimations. Moreover, as shown in Section 16.5.1 [of SRK (2006)] there is a mixture of three grade populations within the mineralized domains. A substantial proportion of the grades representing background values may be spatially associated with the high grade assays. In areas with sparsely spaced data this may lead to over-smoothed block estimates. In view of the above, four types of estimation methods were tested:

“(1) Ordinary kriging (OK) in original space. “(2) Ordinary kriging in unwrinkled space. “(3) Indicator and ordinary kriging of low and high grade populations in original space. “(4) Indicator and ordinary kriging of low and high grade populations in unwrinkled space.

“An additional test using ID2 interpolation was conducted for comparison.”

The comparison of results is provided in Table 17.1.

Table 17.1 Comparison of Results of Five Estimation Methods

Model Million Tonnes Grade (% U3O8) 1 4.82 0.19 2 4.83 0.19 3 4.96 0.18 4 4.92 0.18 ID2 4.93 0.20

SRK considered that Model 2 provided “a better quality estimate with its superior treatment of grade continuities”.

The details of the two block models for the Model 2 estimate are provided in Table 17.2.

Table 17.2 Description of Block Models (Wrinkled and Unwrinkled) for Model 2 (Metres)

Origin, Gauss Kruger Wrinkled Unwrinkled X 281,150 281,150 Y 5,439,050 5,439,050 Z 1,075 12.25 Block Dimensions, m X 15 5 Y 15 5 Z 1.5 1.5 Columns 165 520 Rows 155 495 Levels 420 10 Rotation, counter clockwise 50o 50o

SRK concluded that its Model 2, using ordinary kriging in unwrinkled space and back- transformation into real space, provided the most reasonable estimate of uranium resources for Gurvanbulag. The reader is referred to SRK (2006) for the detailed discussion of Model 2.

SRK (2006) concluded that:

“Comparisons of estimated grades at no cut-off and at elevated cut-offs showed that there was no substantial difference in the metal contents between the different estimates in wrinkled and unwrinkled space, or between the indicator and straight ordinary kriging methods. ID2 on the other hand returned approximately 8% more metal, but was rejected on grounds discussed below. Table 22 [see Table 17.1, above] shows the results of each method at a cut-off grade of 0.07% U3O8. Note that only those blocks that were estimated by all five methods are considered in this comparison.

“Ordinary Kriging (OK) returned slightly better results than the combined indicator and kriging models. These results are also preferable due to the more simple procedure. Although the two OK models are similar in terms of metal contents, Model 2 provides a better quality estimate with its superior treatment of grade continuities. Visual inspections of the two OK models on sections show that in zones of inflection and structural complexity, Model 1 is discontinuous and grades are more restricted (e.g. Figure 22 and Figure 23) [not provided in this report]. In the same areas, Model 2 maintains continuity across inflections, and extends the mineralization further distances across areas of structural complexity. The unwrinkling process in Model 2 resulted in the estimation of blocks that were otherwise left unestimated.”

17.1.3 SRK Mineral Resource Estimate

SRK (2006) reports mineral resources for the Central Zone of the Gurvanbulag deposit at a cut- off grade of 0.07% U3O8, based on a long term uranium price of $47 per pound U3O8 and

SRK’s internal estimate of potential operating costs for underground mining. SRK states that “it is believed [by SRK] to be a reasonable estimate of the potential economic cut-off.” The classified resource estimate is shown in Table 17.3.

Table 17.3 SRK Classified Mineral Resources for the Gurvanbulag Central Zone, Saddle Hills Project, Mongolia, November 3, 2006

Hanging Wall Footwall Totals Thousand % U3O8 Thousand % U3O8 Thousand % U3O8 Thousand Tonnes Tonnes Tonnes Pounds U3O8 Indicated Mineral 570 0.19 2,260 0.22 2,830 0.22 13,633 Resources Inferred Mineral 700 0.13 1,970 0.15 2,670 0.15 8,642 Resources

SRK (2006) describes the classification of the uranium resources as follows:

The ongoing underground sampling and gamma logging program of Western Prospector/Emeelt Mines will provide closely-spaced sampling of areas with higher than average grades and is intended to address the issues noted by SRK, above.

17.2 MINERAL RESERVE ESTIMATION

There has been no estimation of mineral reserves conducted for the Gurvanbulag deposit that is in accordance with the definitions of CIM.

Harper (2005) reported on the historical reserve estimates prepared under the former Soviet classification system. The reader is referred to Harper (2005) for details.

18.0 OTHER RELEVANT DATA AND INFORMATION

The shafts and underground workings at the Gurvanbulag deposit were allowed to flood when the site was abandoned by the Russians in the early-1990s. Western Prospector/Emeelt Mines completed dewatering of the existing shafts and underground workings in 2006.

In addition to exploration, the following work items were also carried out in 2006:

• Completion and lining of settling ponds to accommodate water pumped from the mine.

• Installation of a temporary headframe and hoist at the Gurvanbulag Main shaft.

• Pumping from the Main shaft to dewater all levels down to and including the 260 m level.

• Rehabilitation of the Main shaft to allow access to and cleaning of the 260 m level.

• Establishment of exhaust ventilation at the West shaft.

• Construction of a mine dry and office building to support the underground exploration program.

• Construction of a new 200-person base camp to alleviate pressure on the existing camp.

• Start of construction of a warehouse and vehicle maintenance building at the mine site to provide winter protection during underground exploration.

Access to the 260 m level has been established and the majority of the lateral and inclined development inspected. Systematic sampling of the ore exposed by lateral development is being undertaken to confirm the grade estimates of Russian and Western Prospector/Emeelt Mines diamond drilling.

For the purpose of the preliminary economic assessment of the Gurvanbulag property, Micon retained Malcolm Buck, P.Eng., a mining engineer and Associate of Micon to provide input in developing a suitable mine plan. The geometry of the mineral resources at Gurvanbulag was examined, as presented in SRK (2006). Micon and Mr. Buck also reviewed the updated geology of the deposit in some detail with SRK and with Dr. Harper of Western Prospector. In addition, Eugene Puritch, P.Eng., of P&E, was retained directly by Western Prospector for mine planning. Mr. Puritch provided input to the mine plan prepared for this report. Mr. Buck and Mr. Puritch have visited the Gurvanbulag property on behalf of Western Prospector in January 29-30, 2005 and November 14-18, 2006, respectively. Mr. Puritch’s visit followed the mine dewatering program. He inspected the conditions of the underground working and observed the style of mineralization exposed in headings and cross-cuts.

Micon’s report prepared on behalf of Western Prospector and Emeelt Mines is titled, Preliminary Economic Assessment, Gurvanbulag Uranium Deposit, Mongolia, dated September, 2007.

The geometry of the mineralized zone is essentially tabular, consisting of hanging wall and footwall mineralized zones, generally separated by waste rock, and with thicknesses from less than 1 m to greater than 10 m. The zones dip at approximately 15o to the southeast. It is anticipated that the Gurvanbulag deposit will be mined using underground mining techniques. There are pronounced rolls in the mineralized zones over limited extent, and which the mining method must accommodate. There are indications that, at irregular intervals, the zones may expand and extend in height for tens of metres, though remaining localized in areal extent.

18.1 MINING

• Two 4-m diameter concrete lined shafts, approximately 1 km apart, and both to a depth of 260 m below surface.

• One 6-m diameter concrete lined shaft to a depth of approximately 285 m. This shaft is equipped with a manway and steel piping for dewatering, compressed air and water lines.

• Extensive lateral development on the 260 m level consisting of:

• A main and north haulage crosscut from the 6-m diameter shaft. • Hanging wall haulage drift along 800 m of the mineralized strike length. • Mineralized zones access crosscuts on approximately 50-m centres. • Sill in the mineralized zones. • Footwall ventilation drift connecting the two 4-m diameter shafts. • Boxholes and drawpoints with raises to mineralized zones above.

In general, the shafts and lateral development were found to be in good condition and that little deterioration had taken place since the workings were allowed to flood in the early-1990s.

For the purpose of Micon’s preliminary economic assessment, due to the shallow dip and relatively thin, tabular nature of the deposit, it is proposed that the inclined room and pillar mining method will be applied in the majority of the deposit. In-stope mining will utilize conventional, non-mechanized mining with handheld jacklegs, longtoms and stopers for drilling and slushers for mucking. Mineralized material from ore passes will be removed by load haul dump (LHD) units which will load rail cars. The rail cars will transport mineralized material to dumps near the main shaft. In some areas, small shrinkage or longhole stopes will be used to mine near-vertical lenses of mineralized material.

Figure 18.1 shows the 260 m level existing development and shaft locations.

Figure 18.1 Underground Development and Shaft Locations

Given the mineralization geometry and presently estimated mineral resources, a mine production rate of 1,500 t/d was selected resulting in a mine life of approximately 10 years.

Figure 18.2 shows the 260 m level existing development and shaft locations.

Figure 18.3 shows the 260 m level development and the volumes of mineralized material estimated in mining panels.

Figure 18.2 Plan of 260 Metre Level

Figure 18.3 Plan of 260 Metre Level Showing Estimated Volumes of Mineralized Material

Note: Reef 1 and Reef 2 shown on Figure 17.3 refer to the hanging wall and footwall mineralized zones, respectively.

The 6-m diameter shaft will be the main shaft for transporting rock, men and materials. It has been dewatered and inspected by the mining contractor, Redpath Mongolia LLC. The concrete is dry and in good condition with no visible deterioration. There is an installed manway compartment and piping constructed from steel. The piping is presently being used to provide a dewatering line and compressed air and water to the underground rehabilitation and exploration program, primarily focused on the level with the majority of the underground lateral development (designated 260 m level).

The two 4-m diameter shafts are being used as ventilation shafts and will continue to provide ventilation for the mining operation. Fans are presently installed at the collar of the West shaft to pull exhaust air from the mine. There is no equipment installed in these shafts. They are both sunk to a depth of 260 m and are known to connect to the exploration workings on the 200 m and 260 m levels.

Almost all lateral development on the 260 m level has been found to be in good condition with rail installed in all main drifts and crosscuts. There is also operable compressed air and water piping installed in all openings. The exception is where openings pass through mineralized zones which contain a higher clay content than the wall rocks. As a result the flooding resulted in clay expansion, collapsing mineralized zone rock outwards into the openings resulting in partial to complete caves. Similar effects are manifest in the Russian preparatory stope development above the 260 m level.

Figures 18.4 and 18.5 show the condition of equipment at the exit from the shaft and on the 9900N crosscut on the 260 m level, respectively.

Figure 18.4 Condition of Underground Equipment Following Dewatering Program

E. Puritch, November, 2006

Figure 18.5 View of 9900N Crosscut on 260 m Level

E. Puritch, November, 2006

Mine infrastructure on surface presently consists of a small, temporary headframe servicing the underground exploration program, water treatment settling ponds used to hold water from the underground mine for potential treatment prior to release, a mine dry and mine office, an exploration office building and warehouse/shop facility. All of this infrastructure was installed for the shaft dewatering and underground exploration program and is fully serviceable and designed to be integrated into production facilities.

18.2 MINING SEQUENCE

The mining sequence envisages that mining will start on the 260 m level in the area centred around the 9850N crosscut. This area of mineralization is substantially higher than average grade and would, therefore, provide increased revenues early in the project life. This area was also almost completely developed by the Russians and will facilitate mining. Only the development of raises would be required prior to mining of ore.

Mining will proceed from the 260 m level up dip to the 80 m level. The zones located between the 9600N crosscut and 10400N would be mined first and later zones on this level outside this area would be added. The overall up dip progression of panels would be in an arched shape with the middle panels being mined further up dip than panels near the extremities of the zones. This will mainly help in controlling ventilation flows and maximizing the efficient use of ventilation air sent underground. As mining proceeds past the 160 m level, the mining areas on the 440 m level would be developed and mining there commenced, as required.

Sustaining the 1,500 t/d mining rate requires seven rooms being mined per day. In each panel, 1.5 faces could be mined per day which requires that five panels be mining at one time. A total

of seven panels would be required to be in the mining cycle at any one point in time with the sixth panel being filled and the seventh being prepared to resume mining.

18.3 DEVELOPMENT AND PRODUCTION SCHEDULE

The proposed mining facilities and lateral and raise development will require approximately two years to complete. Mine development will require two development crews, working two 12-h shifts per day.

The proposed development and production schedule for the 1,500 t/d mining rate is shown in Table 18.1.

Table 18.1 Development and Production Schedule for the Gurvanbulag Project

Mineralized Grade Tonnes Production Year Total Block (% U3O8) Between 280,300E and 1 2 3 4 5 6 7 8 9 10 281,000E Level 20 95,000 0.14 95,000 95,000 80 833,000 0.20 100,000 244,000 300,000 189,000 833,000 140 878,000 0.16 54,250 197,000 200,000 243,750 183,000 878,000 200 619,000 0.20 226,750 350,000 42,250 619,000 260 813,000 0.20 547,000 266,000 813,000 320 323,000 0.14 147,000 176,000 323,000 380 348,000 0.19 100,000 248,000 348,000 440 121,000 0.21 32,000 89,000 121,000 Outside 280,300E and

281,000E Level 20 63,000 0.08 63,000 63,000 80 243,000 0.23 154,000 89,000 243,000 140 255,000 0.11 120,000 120,000 15,000 255,000 200 269,000 0.14 65,750 203,250 269,000 260 239,000 0.16 239,000 239,000 320 186,000 0.10 186,000 186,000 380 66,000 0.13 66,000 66,000 Total tonnes 5,351,000 0.17 547,000 547,000 547,000 547,000 547,000 547,000 547,000 547,000 547,000 428,000 5,351,000 Grade 0.20 0.20 0.18 0.16 0.16 0.16 0.17 0.21 0.17 0.13 0.17 (% U3O8) Note: Tonnage by level includes material between it and the level immediately above.

The production schedule is based on a mining recovery of 93% (based on dimensions of pillars left) and dilution of 5% with zero grade.

Material stockpiled on surface comprises approximately 75,000 t at an estimated average grade of 0.07% U3O8. It is planned that this material will be processed during the commissioning phase of the processing plant.

18.4 WASTE ROCK DISPOSAL

All mined waste rock, except that internal to the stoping areas, will be hoisted to surface and sent to a surface waste rock storage area. Waste in stopes will be left in the stopes and incorporated into the backfill.

18.5 MINE SUPPORT FACILITIES

The mine services site will include a paste backfill plant, maintenance shop, offices for mine supervisory, geological, engineering, safelty, health and environment (SHE) and administration staff, the main power substation, warehouse and water treatment facility.

18.6 PASTE BACKFILL

It is planned that all stopes will be backfilled with paste backfill. Tailings from the processing plant will be delivered to a backfill storage tank at the backfill plant. Paste backfill will be prepared and sent to boreholes for distribution to the underground openings. A combination of boreholes and piping on working levels will distribute the paste to stopes being filled.

The conceptual backfill plan is described below.

18.6.1 Backfill Plant

It is proposed that approximately 50% of the tailings stream from the processing plant will be dewatered in a thickener in order to increase the pulp density to greater than 75-80% solids. The thickened tailings will be sent to a mixer and 3-5% cement added to the tailings to create the paste for backfilling. Water will be added to the mixer, as needed, in order to ensure that the paste has the required consistency. The plant will be automated in order to ensure paste consistency and will be monitored from a central control room.

18.6.2 Backfill Distribution

Boreholes will be drilled and used to deliver paste backfill to the main underground distribution level. Boreholes will be 102 mm in diameter and extend to the 80 m level. Piping on this level will distribute the paste to boreholes in order to deliver paste backfill to lower mining levels and over the full strike length of the planned mining operation. Several holes will be required to distribute paste to mining areas over the 600 m to 800 m of the mineralized zones to be developed early in the life of the operation.

Boreholes will be drilled from the 80 m level to the first production level at 260 m. Heavy- wall, 102 mm, steel piping on the level and in the manway accesses will distribute the paste backfill both horizontally and vertically in the stopes.

Boreholes and piping will be completed as required as mining proceeds up and down dip over the life of the operation.

18.7 FLOWSHEET DEVELOPMENT

Melis developed a flowsheet for the processing plant based on the principal design parameters listed in Table 18.2 and the results of the testwork program described above.

Table 18.2 Summary of Key Process Design Parameters

Criterion Units Value Nominal mill feed rate t/y 550,000 Average mill feed grade % U3O8 0.17 Total uranium recovery % 95 Average estimated uranium production lb/y U3O8 1,958,000 kg/y U3O8 888,000

The principal components of the preliminary flowsheet are:

• Crushing and grinding. • Sulphuric acid leaching. • Counter current decantation. • Solvent extraction. • Uranium precipitation. • Yellocake calcination. • Tailings neutralization. • Tailings effluent treatment.

The simplified processing flowsheet is shown in Figure 18.6.

18.8 SITE SELECTION FOR PROCESSING PLANT

Golder was retained by Western Prospector/Emeelt Mines to provide a preliminary assessment of potential sites for the location of the processing plant and tailings management facility (TMF) at Gurvanbulag. These are shown on Figure 18.7. Golder’s report (in draft) is dated January 29, 2007 (Golder (2007)).

For the purpose of the preliminary economic assessment, Western Prospector/Emeelt Mines have selected the location of Alternative 1 for the processing plant which is situated approximately 0.6 km from the Main shaft and some 200 m from the northeast edge of the Option 1 tailings management facility.

Figure 18.6 Simplified Processing Flowsheet

Western Prospector/Emeelt Mines have provisionally selected Option 1 as the preferred location for the tailings management facility. This is based, principally, on the distance from the Main shaft (1.6 km) and in order for the tailings management facility and the mill to be located within the same watershed. A bulk density factor of 1.4 t/m3 was selected in order to determine the capacity requirements for the tailings management facility.

Golder notes that additional work to be carried out for the purposes of feasibility study will include the following:

• Optimization of the site layout and operational plans.

• Geotechnical and hydrogeological investigations to evaluate the suitability of the selected site for development of a tailings management facility.

• Hydrological evaluation of the site.

• Evaluation of tailings handling methods.

• Feasibility level of design for the embankments and other facilities.

Figure 18.7 Alternative Locations of Processing Plant and Tailings Management Facility

18.8.1 Approach to Tailings Disposal

As shown on Figure 18.7, the proposed tailings management facility will be created by building a 760-m long dam across the mouth of the valley provisionally selected as the Option 1 disposal site. The maximum height of the dam is estimated to be approximately 35 m and it will require some 880,000 m3 of fill.

Construction will be rock core covered with sand or gravel. The entire inside of the facility will be lined with HDPE. (Although low permeability clays are used elsewhere in the world to line tailings ponds, suitable clay does not occur within the Gurvanbulag project area.)

Tailings from the processing plant will be slurried with water for pumping to the tailings management facility by pipeline. As noted in Section 16.1.6, testwork undertaken in 2006 indicated that the tailings would be unlikely to generate acid. However, tailings liquid will require treatment for removal of radium and, possibly, for removal of molybdenum also. Sand filters will be required for removal of suspended solids.

It is planned that all water from the tailings management facility will be treated and recirculated back to the processing plant. The facility for treatment of water from the tailings management facility would be used, also, to treat and circulate to the mill water pumped from the underground workings.

For the purpose of Micon’s preliminary economic analysis, it has been estimated that approximately 50% of the total volume of tailings produced in the mill will be used as underground mine backfill. On the basis of the present estimate of mineable material, of 5.4 Mt, it is anticipated that approximately 2.7 Mt of tailings will be stored in the tailings management facility. Golder estimates that the total capacity of the Option 1 site, with the dam built to its maximum height, would be approximately 8.4 Mt. This would provide additional capacity should further mineralization be mined at depth at Gurvanbulag, or to accommodate tailings from the processing of material from satellite deposits that are presently being explored.

18.9 INFRASTRUCTURE

Access to the Saddle Hills area and Gurvanbulag is by unimproved, unpaved road which runs some 130 km north-northwest from Choibalsan, the administrative centre and principal city of Dornod Aimag. Choibalsan is located approximately 570 km east of Ulaanbaatar, the capital city of Mongolia, to which it is connected by partially paved road. Scheduled air service is available between Choibalsan and Ulaanbaatar. A railway line runs north from Choibalsan via Boorj in Russia to connect with the Trans-Siberian Railway line. Rail track remains on the branch line to the Dornod area but is unused, and a graded rail bed runs from there through Mardai and on to Gurvanbulag. There is also a network of roads constructed by the former operators of the Gurvanbulag project between the town site at Mardai and the Gurvanbulag property.

Supplies for construction and operation will arrive by road and rail to Choibalsan and will be transported to Gurvanbulag by road.

The main infrastructure required for the mine will be:

• Main access road from Choibalsan to the mine site. • Warehousing and storage for bulk materials in Choibalsan. • Site roads. • Powerline from Choibalsan to the mine site. • Surface maintenance shop. • Technical services and administration office building. • Warehouse and laydown yard. • Camp and recreational facilities. • Heating plant. • Water supply system. • Landfill site. • Sewage disposal site. • Rail head receiving facility.

The security fence which currently surrounds the project area will be enlarged to encompass all mining and processing facilities, including the tailings management facility. Site security will be provided under contract with personnel on duty at all times.

A general plan showing the facilities presently on site is provided in Figure 18.8.

18.9.1 Road and Power Line

Permanent power supply will be provided by a new 110-kV line from the power station operated by Dornod Energy System LLC (DES) at Choibalsan. The power line is being constructed under a joint venture agreement executed by Emeelt Mines and XinXin which is developing its Ulaan base metal mine 7 km to the east of the Gurvanbulag mine site. In addition, power could be supplied to the soum communities of Bayandun, Dashbalbar and Gurvanzagal, none of which presently is connected to the electrical grid. The feasibility study for the power line was undertaken by Energy International Ltd.

Figure 18.8 General Site Plan

18.9.2 Site Services

Site facilities will meet the particular requirements of the Gurvanbulag uranium project. The support service facilities will be located in proximity to the Main shaft headframe and hoist house and will include the surface maintenance shop, technical services and administration office building and warehouse/laydown yard.

Satellite links presently provide telephone and broad band data services. It is proposed that a fibre-optic cable will be laid along the route of the power line from Choibalsan in order to provide adequate telecommunications capacity for the operation.

The accommodation camp is supplied with non-potable water from a dedicated well. Bottled potable water is used for cooking and drinking.

An on-site coal-fired unit will provide hot water and heat for mine air and for buildings. It is anticipated that this unit will be purchased and shipped from a Russian manufacturer.

Medical services are provided under contract by SOS International through SOS Medica Mongolia LLC. Two full-time doctors are located at site. An ambulance has been provided and a helicopter pad has been constructed for emergency use.

Clean solid waste will continue to be disposed of in an existing permitted disposal site. Sewage will continue to be disposed of in approved septic systems. Hazardous wastes will be disposed of in accordance with industry best practice and in compliance with applicable regulations.

18.9.3 Materials Shipment

Emeelt Mines will acquire warehousing and storage areas in Choibalsan in order to undertake transhipment of bulk and other materials between rail and truck. For incoming materials, it is anticipated that shipments will be consolidated into containers for ease of handling and security. Truck haulage from Choibalsan to Gurvanbulag will be undertaken by contractors who will be required to conform to speed limits and maintenance schedules set by Western Prospector/Emeelt Mines. It is proposed that product from the processing plant will be in sealed 205-L steel vessels and shipped from site in secured ocean containers.

18.10 ENVIRONMENTAL, SOCIO-ECONOMIC CONDITIONS AND PERMITTING

The information presented in Section 18.10 has been provided by Western Prospector/Emeelt Mines.

Mineral exploration and development of a uranium mining project in Mongolia are permitted within the framework of the following three principal areas of legislation:

• Minerals Law of Mongolia, 2006. • Law of Mongolia on Radiation Protection and Safety, 2001. • Law of Environmental Impact Assessment, 1998.

It is understood that Emeelt Mines/Western Prospector provided an environmental protection plan, as required for exploration on the Gurvanbulag licenses, in 2006 and 2007 and that this was approved by the relevant authorities of Sergelen, Dashbalbar and Bayandun soums.

A permit to explore for uranium, as required under the law relating to radiation protection and safety, was received for the licenses held by Western Prospector and Emeelt Mines, and held jointly with Adamas, in 2005. These were renewed in 2006 and 2007. Periodic inspections are made by staff of the Nuclear Regulatory Authority and, as reported by Western Prospector/Emeelt Mines, none of has resulted in suspensions or censoring.

18.10.1 Environmental Baseline Study

An environmental baseline study has been prepared by EcoTrade LLC on behalf of Western Prospector Mongolia (EcoTrade (2005)) in order to provide data on the geography, geology, hydrology, hydrogeology, soil, climate, flora and fauna, and the socio-economic conditions of the area. Subsequently, EcoTrade has continued to undertake monitoring on an approximately quarterly basis.

The environmental baseline study covers geography, geology, hydrology, hydrogeology, soil, climate, flora and fauna and the socio-economic conditions in the project area. Previously disturbed areas were documented and a comparison was undertaken of radiation protection laws, regulations and standards in Mongolia, Australia and Canada. Field studies were carried out in April, June, July and September, 2005 for the environmental baseline report.

The baseline study meets the requirements for screening under Mongolian environmental impact assessment procedures. Emeelt Mines notes that submission of the baseline study, together with a Project Proposal, which describes the development of the Gurvanbulag project, will meet the screening criteria to establish the need for an environmental impact assessment (EIA) and the guidelines for its completion.

Reports on the environmental impact assessment and an environmental protection plan and monitoring program for the power line from Choibalsan have been prepared by Ecos LLC (see Ecos LLC (2007a) and (2007b)). At the time of writing, both reports are available only in Mongolian.

18.10.2 Previously Disturbed Areas

There are significant areas of previously disturbed land resulting from Russian exploration and mining activities and supporting infrastructure including pipelines, abandoned railway lines, and the abandoned town site of Mardai.

At the request of Western Prospector/Emeelt Mines, these previously disturbed areas were documented by EcoTrade in four separate reports. The disturbed areas are also described in the EcoTrade baseline study.

During 2005, Western Prospector/Emeelt Mines cleared the majority of unsightly and unsafe remnants of abandoned structures in the vicinity of Gurvanbulag.

In 2006, Western Prospector continued work to identify, monitor and remediate areas contaminated by low-level radiation in the Gurvanbulag area. The Mongolian Regulatory Authority (RA) carried out a parallel program of monitoring. Results were similar.

Areas which required particular attention were excavated and the contaminated material stockpiled within the double-fenced, approved facility for radioactive materials storage established at the mine site. The excavated areas were then backfilled with checked clean fill and the areas resurveyed to ensure their compliance. The radioactive materials storage site was established and approved by the RA in 2006 and contains primarily the mineralized rock extracted from the underground development by the former workers.

The radiation levels measured in undisturbed areas are only slightly elevated, do not represent a threat to health or safety and, therefore, are of no particular concern. The highest levels, recorded at the Main shaft dump, are of the stockpile of low-grade mineralized material noted above. The area has been fenced and provided with signs in order to prevent inadvertent entry. It is intended that this material will be processed for recovery of uranium when the processing plant becomes operational.

18.10.3 Dewatering

Mine dewatering activities may be carried out under an amendment to the environmental protection plan filed to support an exploration license.

EcoTrade LLC was retained to prepare an environmental impact assessment and environmental protection plan for dewatering at Gurvanbulag in July, 2005. The plan was approved in October, 2005.

Provisions under the environmental protection plan relate to:

• Discharge and treatment of water from underground workings. • Monitoring of discharge water quality. • Measurement of soil contamination through the flow channel. • Monitoring of vegetation and fauna near the flow channel and adjacent areas.

All of the provisions were met during the dewatering program.

In order to treat water pumped from underground, a system of settling ponds, lined with HDPE, was constructed. The first pond, with a volume of 4,200 m3, provides for settling of solids and the second, volume 8,300 m3, provides for treatment. A cascade system at the entry to the first pond and between the two ponds aerates the water and stabilizes the pH. Any dissolved iron forms iron hydroxide which precipitates and settles in the ponds. Aeration also lowers the dissolved radon gas content.

Final discharge from the second pond is into a shallow valley which runs to an accumulation basin approximately 9 km from the discharge point. During pumping in 2006, all discharged water went to ground within 3 km of the discharge point.

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Water pumped from the mine to the ponds contained low suspended solids.

A small plastic-lined pad was also constructed to provide temporary storage for contaminated materials brought up from underground prior to decontamination or disposal. The pad was not used during 2006, or to date in 2007.

18.10.4 Groundwater

In accordance with the Environmental Management Plan, a network of monitoring wells was set up in order to measure the impact of the mine dewatering program on groundwater and to monitor natural seasonal variations in the groundwater regime. Many of these were vertical water wells drilled during Russian exploration work and were able to be used by Emeelt Mines for the purposes of groundwater monitoring.

The results show that the water levels in wells 368, 379 369 and 370 are moderately affected by dewatering while others (wells 371, 305 and 374) show little drawdown and the remaining 13 wells show unchanged water levels. Also, there was no correlation between well drawdown and the distance from the point of pumping in the main shaft. This indicates that the aquifer is likely to be fracture-controlled with drawdown being greatest when the well intersects a water-bearing fracture.

There was a relationship between water drawdown and depth with the deeper wells being affected most. This may indicate that a deeper, confined aquifer is overlain by a shallow, unconfined aquifer.

Monitoring of groundwater during mine dewatering indicates that the dewatering program is drawing from the deep, unconfined aquifer but has little effect on the shallow, unconfined aquifer. Water that is used by animal herders or for agriculture is typically drawn from shallow wells dug by hand, or from surface ponds or streams that are fed by the unconfined aquifer. On the basis of information gained from the monitoring program, these shallow water sources are not affected by mine dewatering.

18.10.5 Socio-economic Considerations

A Community Relations Officer was engaged in August, 2006 with responsibility for liaison with communities and official representatives. During 2006, three information meetings were conducted in communities closest to Gurvanbulag. These meetings were chaired by the local soum governors and representatives of Emeelt Mines were available to respond to questions and discussion.

An advisory committee of representatives from each of the communities, Sergelen, Dashbalbar and Bayandun, has been set up. The committee has met quarterly since the autumn of 2006 at Gurvanbulag to tour the operation, to receive progress reports and to bring forward any concerns or questions from their constituent communities.

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In 2006, herders within 25 km of the project were interviewed in order to determine the size of the herds, the use of seasonal pasture, water supplies and the availability of skills that may be employed by Emeelt Mines.

In 2007, a survey to gain information on available skills was carried out in the local soums.

18.10.6 Safety, Health and Environment

The Radiation Protection Program is carried out under the general direction of the Radiation Officer who is assisted by staff technicians and qualified consultants. The program includes a range of monitoring and training components and has been designed to meet both Mongolian and international standards and best practices.

In 2005, in conjunction with the Mongolian RA, a program was introduced whereby all workers wore personal dosimeters in order to measure exposure to radiation. The scope of this program was reduced when it was found that the majority of measurements did not exceed background and dosimeters were used only by those workers likely to be exposed to elevated radiation levels.

Following dewatering of the underground workings, an appropriate radiation program was implemented in advance of the commencement of the underground sampling work. Underground work places are monitored daily and personnel access and ventilation adjusted accordingly. Training in radiation protection is provided by qualified Emeelt Mines personnel and expatriate experts.

A safety manual for workers, supervisors and visitors to site was completed in May, 2005 in both English and Mongolian. The manual, which has been revised periodically, provides safety rules, guidelines and directions to be followed by all personnel.

Having consulted with the Mongolian authorities, mine regulations applicable in the province of Ontario, Canada, were implemented at Gurvanbulag to provide the framework for underground safety at the start of the underground sampling program.

A formal closure plan for the proposed Gurvanbulag mining and processing operation has yet to be developed. The objectives for the site at closure will be the following:

• Minimal level of ongoing emissions of radiation above background levels.

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• Removal of all structures to ground level. • Installation of a suitable cover for the tailings disposal areas. • Underground mine workings to be allowed to flood. • Recontouring of the site to blend with the surrounding terrain.

In contrast to other uranium mining operations, it is anticipated that tailings from an operation at Gurvanbulag will contain minimal sulphide minerals, will be stable chemically and will not produce sulphuric acid. The objective at Gurvanbulag will be to store the tailings in a secure facility which will not fail, and to minimize release of any hazardous material from the tailings. It is anticipated that this will be accomplished by covering the tailings and by engineering the cover so that water does not accumulate behind the dam of the tailings pond.

18.11 THE MARKET FOR URANIUM

Colorado Nuclear, Inc. (CNI) was retained by Western Prospector to provide an analysis of the international uranium market. Its report, Global Uranium Market – Demand, Supply and Price Forecast, dated May, 2007, provides data on the uranium industry and a projection of long term uranium prices for use in the economic analysis of the Gurvanbulag project. (See Colorado Nuclear, Inc. (2007). The following short description of the market for uranium is based on the CNI report.

World uranium requirements to fuel reactors are shown in Table 18.3.

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Table 18.3 World Uranium Requirements, 2005-2030 (Million pounds U3O8)

Year Reference Case Upper Case Lower Case 2005 170.0 170.0 170.0 2006 170.2 170.7 169.0 2007 173.0 173.6 171.6 2008 173.2 183.6 172.9 2009 178.2 190.1 172.4 2010 186.0 193.1 176.5 2011 188.7 201.9 176.4 2012 191.1 208.6 178.3 2013 197.1 212.4 178.3 2014 201.6 221.5 175.3 2015 202.8 230.1 178.1 2020 220.3 267.5 173.0 2025 260.4 339.6 157.1 2030 288.0 413.9 137.6 Total 5,729.1 6,878.5 4,369.4 World Nuclear Association, 2005

World natural uranium production rose slightly during 2005 reaching 108.9 Mlb U3O8 compared to the aggregate global output of 105.7 Mlb U3O8 reported for 2004.

Uranium production in Australia, Canada, Kazakhstan, Namibia, Russia, the United States and Uzbekistan increased during 2005, with Niger and South Africa showing a slight decline.

As has been the case for the past several years, a total of 17 countries had uranium production take place during 2005. The top 12 countries produced 99.0% of the total output (107.9 Mlb U3O8) with the top eight countries contributing 101.0 Mlb U3O8 , or 93 % of the aggregate. See Table 18.4.

Table 18.4 World Natural Uranium Production by Country, 2005

Country Tonnes U Million pounds U3O8 Share of Total (%) Canada 11,628 30.2 27.8 Australia 9,522 24.8 22.7 Kazakhstan 4,357 11.3 10.4 Russia 3,431 8.9 8.2 Namibia 3,147 8.2 7.5 Niger 3,093 8.0 7.4 Uzbekistan 2,300 6.0 5.5 United States 1,039 2.7 2.5 Ukraine 800 2.1 1.9 China 750 2.0 1.8 South Africa 673 1.8 1.6 Czech Republic 408 1.1 1.0 Subtotal 41,490 107.9 99.0 Others 1 380 1.0 1.0 Total 41,870 108.9 100.0 1 Includes Brazil, Germany, India, Pakistan, Romania. World Nuclear Association

Based upon current and projected uranium market conditions which will continue to be supply constrained, CNI provided the following projection of uranium market prices,

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specifically for the Gurvanbulag deposit. Beginning in 2015, the long-term sustainable uranium price was estimated by CNI at $80.00/lb U3O8, expressed in constant, 2007 dollars.

Table 18.5 Long-term Price Projection ($ per pound U3O8, constant 2007 $)

2007 130.00 2008 150.00 2009 160.00 2010 140.00 2011 120.00 2012 110.00 2013 100.00 2014 90.00

Natural uranium concentrates planned to be produced from the Gurvanbulag project may be sold at world market prices either under near-term delivery contracts (spot market) or under multi-year delivery agreements (term contracts). It is understood that Western Prospector/Emeelt Mines do not have have uranium sales agreements in hand.

18.12 CAPITAL COST ESTIMATE

Total capital costs for the project, in constant United States dollars of mid-2007 value, have been estimated, as detailed in Table 18.6, at $269 M. Of this amount, some $229 M, including a working capital of some $12 M, will be required as pre-production capital to the point where the project starts to generate a positive operating margin. These estimates, unless otherwise noted, include a contingency of 15%. Capital costs have been estimated to a level of detail appropriate for a preliminary economic assessment, in this case, to an overall level of +/-25%.

Table 18.6 Summary of Capital Expenditure (Thousand dollars)

Item Pre-production Post-production Total Life of Mine Exploration and feasibility study 8,000 - 8,000 Mine development 11,644 41,668 53,312 Mining equipment 6,898 1,824 8,722 Underground fixed equipment 18,035 2,439 20,474 Processing plant 126,036 600 126,636 Infrastructure and services 30,295 600 30,895 EPCM 15,947 - 15,947 Reclamation and closure - 5,000 5,000 Working capital 11,822 (11,822) - Total 228,677 40,309 268,986

18.13 OPERATING COST ESTIMATE

The total unit operating cost for the project is estimated at $86.45/t of material processed, as detailed in Table 18.7. Operating costs have been estimated to a level of detail appropriate for a preliminary economic assessment, in this case, to an overall level of +/-25%.

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Table 18.7 Summary of Unit Operating Costs

Item $/t Mining 36.99 Mine management manpower 1.43 Processing plant 34.50 Surface services manpower 1.57 General services and administration 11.96 Total 86.45

18.14 PROJECT SCHEDULE AND MANPOWER

A schematic diagram of the project schedule is presented in Figure 18.9. It is anticipated that the majority of work carried out in year -3 will consist of exploration, detailed engineering and procurement of equipment with long lead time. A full feasibility study is also planned to be completed during year -3, as is work related to securing the operating permit. Construction leading to production will be undertaken in years -2 and -1.

18.15 ECONOMIC EVALUATION

Based on the indicated and inferred mineral resources estimated by SRK for the Central Zone of the Gurvanbulag deposit and on the results of the underground dewatering program completed at the end of 2006, Micon has developed a plan to mine and process an average of 1,500 t/d of mineralized material that results in mining all of the presently estimated 5.4 Mt of resources over a project life of about 10 years.

The economic evaluation is treated on a full project basis, i.e., assuming 100% equity financing and, for the base case, a uranium price of $185/kg U3O8 (equivalent to $84/lb U3O8) has been assumed.

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Figure 18.9 Potential Project Development Schedule

Activity Year -3 Year -2 Year -1 Year 1 M -12 M -11 M -10 M -9 M -8 M -7 M -6 M -5 M -4 M -3 M -2 M -1 M -12 M-11 M -10 M -9 M -8 M -7 M -6 M -5 M -4 M -3 M -2 M -1 M -12 M-11 M -10 M -9 M -8 M -7 M -6 M -5 M -4 M -3 M -2 M -1

Mine

3 Headframe and Hoist Final Detailed Engineering

4 Headframe Purchase and Fabrication

5 Hoist Refurbishing and Delivery

6 Long Lead Time Equipment Procurement and Delivery

7 Detailed Shaft & Shaft Facilities Design

8 Shaft Rehabilitation

9 Shaft Underground Facilities

10 Detailed Mine Design

11 Mine Development

12 Ore Mining

Processing Plant

13 Metallurgical Testwork

17 Processing Plant Detailed Design

18 Long Lead Time Equipment Procurement and Delivery

19 Processing Plant Construction

20 Tailings Storage Construction

Infrastructure

21 Infrastructure Detailed Design

22 Powerline & Surface Power Distribution

23 Access Road

107 24 Surface Buildings

25 Other Facilities

Summary results of the preliminary economic assessment are provided in Table 18.8.

Table 18.8 Summary of Results of Preliminary Economic Assessment of the Gurvanbulag Deposit

Item Unit Value Pre-production capital cost $M 228.7 Life of mine capital cost $M 269.0 Operating costs $M 462.6 Cash operating cost $/t milled 86.45

Cash operating cost $/kg U3O8 produced 52.06 Total uranium production thousand kg U3O8 8,886 Total project cash flow $M 830.1 Income tax payable $M 187.4 Royalty payable $M 82.21 Project cash flow (after tax) $M 642.7 NPV @ 10%/y discount rate $M 241.5 Internal rate of return % 35 1 Deducted from revenue.

Among the factors assumed in the development of the cash flow model are:

Based on these assumptions and analyses, the project cash flow yields an IRR of 35% after tax. The NPV at a discount rate of 10%/y is $241 M.

The results of the base case cash flow for preliminary economic assessment of the Gurvanbulag deposit are provided in Table 18.9.

Table 18.10 shows the sensitivity of the project cash flow to uranium price, mined grade, operating costs and capital costs. As may be expected, factors which directly affect revenue, i.e., the price of uranium and mined grade, are the most sensitive items, followed in turn by capital and then operating costs. The results on IRR and net present value of varying the grade of uranium by 10% and 25% are the same as for the price of uranium.

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Table 18.9 Gurvanbulag Preliminary Economic Analysis

FOR INTERNAL USE OF WESTERN PROSPECTOR GROUP ONLY

RESOURCES

YEAR Year -3 Year -2 Year -1 Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10 Total

PRODUCTION Ore Production (000t) 0 0 0 547 547 547 547 547 547 547 547 547 428 5,351 Ore Grade (%U3O8) 0.20 0.20 0.18 0.16 0.16 0.16 0.17 0.21 0.17 0.13 0.17 Contained U3O8 (000Kg) 1,094 1,094 985 875 875 875 930 1,149 930 556 9,363 Recovered U3O8 (000Kg) 94.90% 1,038 1,038 934 831 831 831 882 1,090 882 528 8,886

REVENUES ($000) Total Revenue @$185.00/Kg 185 192,068 192,068 172,861 153,654 153,654 153,654 163,258 201,672 163,258 97,684 1,643,833 Less Royalties @5.0% 5.00% 9,603 9,603 8,643 7,683 7,683 7,683 8,163 10,084 8,163 4,884 82,192 Net Revenue 182,465 182,465 164,218 145,972 145,972 145,972 155,095 191,588 155,095 92,800 1,561,641

OPERATING COSTS $/t Mining 36.99 36.99 1 20,234 20,234 20,234 20,234 20,234 20,234 20,234 20,234 20,234 15,832 197,933 Processing 34.50 34.5 18,872 18,872 18,872 18,872 18,872 18,872 18,872 18,872 18,872 14,766 184,610 109 Infrastructure & Services 1.57 1.57 859 859 859 859 859 859 859 859 859 672 8,401 Mining Manpower 1.43 1.43 782 782 782 782 782 782 782 782 782 612 7,652 Gen. & Admin 11.96 11.96 6,542 6,542 6,542 6,542 6,542 6,542 6,542 6,542 6,542 5,119 63,998 Total Operating Costs 86.45 86.45 47,288 47,288 47,288 47,288 47,288 47,288 47,288 47,288 47,288 37,001 462,594

OPERATING CASH FLOW ($000) 135,177 135,177 116,930 98,684 98,684 98,684 107,807 144,300 107,807 55,800 1,099,047

CAPITAL COSTS ($000) Exploration 20,100 7,000 7,000 Feasibility Study 1,000 1,000 Mine Development 6,380 5,264 776 4,704 12,052 1,680 7,005 4,917 6,900 3,634 0 53,312 Mining Equipment 1,840 5,058 0 0 81 6 696 1,006 6 6 23 0 8,722 U/G Fixed Equipment 7,717 10,232 86 288 132 132 937 288 288 288 86 0 20,474 Processing Plant 2,000 77,536 46,400 100 100 100 100 100 100 100 0 0 0 126,636 Infrastructure & Services 7,300 5,300 13,823 11,072 100 100 100 100 100 100 100 0 0 0 30,895 EPCM 8,712 7,235 15,947 Reclamation & Closure 5000 5,000 00 Working Capital 11,822 -11,822 0

Total Capital Costs 1 27,400 15,300 109,628 86,377 17,372 1,264 5,117 12,390 3,513 8,499 5,411 7,194 3,743 -6,822 268,986

PROJECT CASH FLOW ($000) -15,300 -109,628 -86,377 117,805 133,913 111,813 86,294 95,171 90,185 102,396 137,106 104,064 62,622 830,061 Less: Capital Write Off 42,261 42,261 42,261 42,261 42,261 211,305 Less: Tax Loss Carry Forward 53,031 27,400 16,086 16,086 16,086 16,086 16,086 80,431 Taxable Income 59,457 75,565 53,466 27,946 36,823 90,185 102,396 137,106 104,064 62,622 749,630 Taxes @ 25% 25% 14,864 18,891 13,366 6,987 9,206 22,546 25,599 34,276 26,016 15,655 187,408 Net Cash Flow -15,300 -109,628 -86,377 102,940 115,021 98,447 79,307 85,965 67,638 76,797 102,829 78,048 46,966 642,654 Cumulative Cash Flow -15,300 -124,928 -211,305 -108,365 6,656 105,103 184,410 270,375 338,013 414,810 517,639 595,687 642,654

IRR 35% NPV @8% $293,752 NPV @10% $241,500

Table 18.10 Results of Sensitivity Analysis

The results of full feasibility study of the Gurvanbulag project are likely to differ from the results of Micon’s preliminary economic assessment.

Micon is not aware of any additional information or explanation necessary in order to make this technical report understandable and not misleading.

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19.0 INTERPRETATION AND CONCLUSIONS

Based on the mineral resources estimated by SRK for the Central Zone of the Gurvanbulag deposit, on the results of metallurgical testwork completed in December, 2006 and on the results of the underground dewatering program completed at the end of 2006, Micon has prepared a preliminary economic assessment of the Gurvanbulag property which envisages the mining and processing of an average of 1,500 t/d of mineralized material that results in mining all of the presently estimated 5.4 Mt of resources over a project life of about 10 years.

The base case discounted cash flow analysis of the Gurvanbulag project yields an IRR of 35% after tax. In common with virtually all mining developments, the project cash flow is most sensitive to revenue, in this case, to the price of uranium and grade of the deposit. The result of reducing either the price of uranium or the mined grade by 25% is an IRR of 22% which supports the robustness of the project at this level of preliminary economic assessment.

As noted above, that the results of full feasibility study of the Gurvanbulag project are likely to differ from the results of Micon’s preliminary economic assessment.

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20.0 RECOMMENDATIONS

On the basis of the results of the preliminary economic analysis, it is recommended that Western Prospector/Emeelt Mines proceeds with work on feasibility studies for the Gurvanbulag project.

It is recommended, also, that:

• The results of exploration through the second half of 2006 and 2007 are compiled into an updated estimate of mineral resources.

• Work is continued on development of the main access road and the power line and distribution facilities are completed.

• Work is continued on environmental baseline studies and community development programs.

• Work is continued towards securing the permits that will be required for construction and development.

• The assessment of the uranium market is updated to take account of recent movements in the spot price for uranium and its potential impact long term contract prices that will be applicable to the Gurvanbulag project.

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21.0 SIGNATURE PAGE

The present Independent Technical Report discloses the results of a preliminary economic assessment of the Gurvanbulag uranium deposit which was completed on October 15, 2007.

The effective date of the present Independent Technical Report is November 27, 2007.

Signed:

“Malcolm Buck”

Malcolm Buck, P.Eng., Associate Mining Engineer, Micon International Limited

“Bruce C. Fielder”

Bruce C. Fielder, P.Eng., Senior Processing Engineer, Melis Engineering Ltd.

“Marek Nowak”

Marek Nowak, MASc., P.Eng., Principal Geostatistician, SRK Consulting (Canada) Inc.

“Eugene J. Puritch”

Eugene Puritch, P.Eng., Principal, P&E Mining Consultants Inc.

“Jane Spooner”

Jane Spooner, P.Geo., Principal, Micon International Limited

“Mani M. Verma”

Mani M. Verma, P.Eng., Associate Mining Engineer, Micon International Limited

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22.0 REFERENCES

Colorado Nuclear, Inc., 2007: Global Uranium Market – Demand, Supply and Price Forecast, Prepared for Saddle Hills Uranium Project – Gurvanbulag Deposit (Western Prospector Group Ltd.), May, 2007.

Ecos LLC, 2007a: Choibalsan – Emeelt – Xinxin 110kV Transmission Line Project, Environment Protection Plan and Monitoring Program, 2007 (in Mongolian).

Ecos LLC, 2007b: Report on the Environmental Impact Assessment for Choibalsan – Emeelt – Xinxin 110kV Transmission Line, for Emeelt Mines LLC, 2007 (in Mongolian).

EcoTrade LLC, 2005: Saddle Hills Uranium Project, Environmental Baseline Study Report on Gurvanbulag Uranium Deposit Area within Emeelt Hills, Bayandun Soum, Dornod Aimag, Mongolia, 2005.

Golder Associates Ltd., 2007: Technical Memorandum, Preliminary Site Selection, Gurvanbulag Tailings Management Facility, Mongolia, January 29, 2007 (in draft).

Harper, Gerald, 2005: Report on Saddle Hills area uranium prospects, Exploration Licenses 3367X, 4846X, 4969X, 4970X, 4974X, 6995X, 7023X, 7405X, 7685X and 9283X, Dornod Aimag, eastern Mongolia, for Western Prospector Group Ltd., May 4, 2005.

Melis Engineering Ltd., 2007a: Emeelt Mines LLC, and Western Prospector Group Ltd., Gurvanbulag Deposit, Saddle Hills Uranium District, Mongolia, Summary of Metallurgy, dated February 14, 2007.

Melis Engineering Ltd., 2007b: Process Plant Capital and Operating Cost Estimates for Gurvanbulag Deposit, Saddle Hills Uranium District, Mongolia – Rev. 1, Memorandum, August 2, 2007.

Micon International Limited, 2007: Western Prospector Group Ltd. and Emeelt Mines LLC, Preliminary Economic Assessment, Gurvanbulag Uranium Deposit, Mongolia, September, 2007.

SRK, 2006: Independent Technical Report and Resource Estimate for the Gurvanbulag Deposit; Saddle Hills Uranium Project, Mongolia, November 17, 2006.

Steffen, Robertson and Kirsten, et al., 1987, Canadian Uranium Mill Waste Disposal Technology, Canada Centre for Mineral and Energy Technology.

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CERTIFICATE

Malcolm Buck

As an author of this report, I, Malcolm Buck do hereby certify that: 1. I am employed as an Associate Mining Engineer, and carried out this assignment for; 2. Micon International Limited, Suite 900, 390 Bay Street, Toronto, Ontario M5H 2Y2, tel (416) 362 5135, fax (416) 362 5763. 3. I hold the following academic qualifications: B.Eng. (Mining Engineering), Technical University of Nova Scotia, Halifax, N.S., 1983. Masters of Engineering (Mineral Economics), McGill University, Montreal, P.Q., 1986. 4. I am registered as a Professional Engineer with Professional Engineers Ontario with registration number 5881503. In addition, I am a member of the National and Toronto Branch of the Canadian Institute of Mining, Metallurgy and Petroleum. 5. I have worked as a mining engineer in the minerals industry for over 24 years. 6. I am familiar with NI 43-101 and, by reason of education, experience and professional registration, I fulfill the requirements of a Qualified Person as defined in NI 43-101. My work experience includes undertaking numerous mine feasibility studies (including uranium mines), as well as performing operating and engineering functions at operating mines in Canada and elsewhere in the world. 7. I visited the Gurvanbulag uranium property on January 29-30, 2005. 8. I am responsible for Sections 18.1, 18.2, 18.3, 18.4, 18.5 and 18.6 of the technical report entitled, “Independent Technical Report on the Results of a Preliminary Economic Assessment, Gurvanbulag Uranium Deposit, Mongolia”, that deal with mining plans and mining capital and operating cost estimates; I also contributed to the preparation of Section 18.9, Infrastructure. 9. Previously, I participated in the preparation of an internal report entitled, “Underground Dewatering Program and Cost Estimate, Gurvanbulag Deposit”, November, 2005, and in the preparation of an internal economic assessment of the Gurvanbulag property, September, 2005. 10. I am independent of the issuer under Section 1.4 of NI 43-101.. 11. As of the date of this certificate, to the best of my knowledge, information and belief, the technical report contains all scientific and technical information that is required to be disclosed to make the technical report not misleading.

Dated this November 27, 2007

“Malcolm Buck”

“Malcolm Buck” Malcolm Buck P.Eng.

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CERTIFICATE OF QUALIFIED PERSON Bruce C. Fielder, P.Eng. Principle Process Engineer, Melis Engineering Ltd. Suite 100, 2366 Avenue C North, Saskatoon, Sk, S7L 5X5 Tel: (306) 652-4084 Fax: (306)653-3779 Email: [email protected]

I, Bruce C. Fielder, am a Registered Professional Engineer in the Province of Saskatchewan, Registration No. 10309. I am employed by Melis Engineering Ltd., 2366 Avenue C North, Suite 100, Saskatoon, Saskatchewan, S7H 4N1, as Principal Process Engineer. I graduated from the University of Alberta, Edmonton, Ab., with a BSc. Degree in Metallurgical Engineering in 1981 and am a member of the Canadian Institute of Mining Metallurgy and Petroleum. I have practiced my profession continuously since 1981 and have been involved in: test work supervision, preparation of process audits, scoping, pre-feasibility, and feasibility level studies for copper, gold, molybdenum, zinc and uranium properties in Canada, United States, Mongolia and China. I am currently a Consulting Engineer and have been so since 2001. I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101. I served as the Qualified Person for Sections 16 and 18.7 of the Independent Technical Report on the Results of a Preliminary Economic Assessment, Gurvanbulag Uranium Deposit, Mongolia (Independent Technical Report), relating to metallurgical testwork and process design. The metallurgical testwork was completed at SGS Lakefield Research Limited under my direction. I have visited the Gurvanbulag property on October 13th-25th, 2006 and June 15th-20th, 2007. As of the date of this certificate to my knowledge, information and belief, the Independent Technical Report contains all scientific and technical information that is required to be disclosed to make this report not misleading. I am independent of Western Prospector Group Ltd. and Emeelt Mines LLC in accordance with the application of Section 1.4 of National Instrument 43-101. I have read National Instrument 43-101 and certify that the portions of the Independent Technical Report for which I served as a Qualified Person have been prepared in compliance with that Instrument. Dated 26 November, 2007.

“Bruce C. Fielder” Bruce C. Fielder, P.Eng.

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To accompany the Independent Technical Report by Micon International on the Results of a Preliminary Economic Assessment, Gurvanbulag Uranium Deposit, Mongolia, dated November 27, 2007

I, Marek Nowak, hereby certify that:

1. I reside at 1307 Brunette Ave. Coquitlam, BC V3K 1G6, Canada;

2. I am a Principal Geostatistician with the firm of SRK Consulting (Canada) Inc. (SRK) with an office at Suite 2200-1066 West Hastings Street, Vancouver BC, Canada;

3. I have a Master of Science degree from the University of Mining and Metallurgy, Cracow, Poland, and a Master of Science degree from the University of British Columbia, Vancouver, Canada;

4. I am a member in good standing of The Association of Professional Engineers and Geoscientists of the Province of British Columbia, Canada, Reg. No. 16985;

5. I have over 25 years of experience in the mining industry. I specialize in natural resource evaluations;

6. I have not received, nor do I expect to receive, any interest, directly or indirectly, in the subject property or securities of Western Prospector Group Ltd.;

7. I have read National Instrument 43-101 and Form 43-101F1 and I am a Qualified Person for the purpose of NI 43-101. My work experience includes 16 years as a geostatistician, five years as a geologist and four years as a mining engineer.

8. I am a co-author of the Independent Technical Report and Resource Estimate for the Gurvanbulag Deposit, November 17, 2006.

9. I am responsible for SRK resource estimates described in Section 17.1 of this report. The resource estimates are described in detail in Section 16 of the Independent Technical Report and Resource Estimate for the Gurvanbulag Deposit, November 17, 2006.

Dated this 27th day of November, 2007

Marek Nowak, MASc., P.Eng.

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CERTIFICATE OF AUTHOR

EUGENE J. PURITCH, P. ENG.

I, Eugene J. Puritch, P. Eng., President of P&E Mining Consultants Inc., practicing from 2 County Court Blvd., Suite 202, Brampton, Ontario, L6W 3W8, do hereby certify that:

1. I am a graduate of The Haileybury School of Mines, with a Technologist Diploma in Mining, as well as obtaining an additional year of undergraduate education in Mine Engineering at Queen’s University. In addition, I have also met the Professional Engineers of Ontario Academic Requirement Committee’s Examination requirement for Bachelor’s Degree in Engineering Equivalency. I have practiced my profession continuously since 1978. My summarized career experience is as follows:

- Mining Technologist - H.B.M.&S. and Inco Ltd. 1978-1980 - Open Pit Mine Engineer – Cassiar Asbestos/Brinco Ltd 1981-1983 - Pit Engineer/Drill & Blast Supervisor – Detour Lake Mine 1984-1986 - Self-Employed Mining Consultant – Timmins Area 1987-1988 - Mine Designer/Resource Estimator – Dynatec/CMD/Bharti 1989-1995 - Self-Employed Mining Consultant/Resource-Reserve Estimator 1995-2004 - President – P&E Mining Consultants Inc. 2004-Present

2. I am a mining consultant currently licensed by the Professional Engineers of Ontario (License No. 100014010) and registered with the Ontario Association of Certified Engineering Technicians and Technologists as a Senior Engineering Technologist. I am also a member of the National and Toronto CIM.

3. I contributed to portions of Section 18 of this technical report titled “Independent Technical Report on the Results of a Preliminary Economic Assessment Gurvanbulag Uranium Deposit, Mongolia” dated November 27, 2007.

4. I visited the Gurvanbulag Uranium Project between Nov 14 to 18, 2006.

5. As of the date of this certificate, to the best of my knowledge, information, and belief, the technical report contains all scientific and technical information that is required to be disclosed to make the technical report not misleading.

6. I am independent of the issuer applying the test in Section 1.4 of NI 43-101.

7. I have had no prior involvement with the Gurvanbulag Uranium Project that is the subject of this Technical Report.

8. I have read NI 43-101 and Form 43-101F1 and the Report has been prepared in compliance therewith.

9. I am a “qualified person” for the purposes of NI 43-101 due to my experience and current affiliation with a professional organization (Professional Engineers of Ontario) as defined in NI 43-101.

DATED this 27th day of November, 2007

“SIGNED AND SEALED” ______Eugene J. Puritch, P.Eng.

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Certificate for Jane Spooner

I, Jane Spooner, do hereby certify that:

1. I am employed by, and carried out this assignment for, Micon International Limited, Suite 900, 390 Bay Street, Toronto, Ontario M5H 2Y2, tel. (416) 362-5135, fax (416) 362-5763, e-mail jspooner@micon- international.com;

2. I hold the following academic qualifications: B.Sc. Hons. Geology University of Manchester 1972 M.Sc. Environmental Resources University of Salford 1973

3. I am a registered Professional Geoscientist with the Association of Professional Geoscientists of Ontario (Registration Number 0990); as well, I am a member in good standing of other technical associations and societies, including: The Prospectors and Developers Association of Canada Canadian Institute of Mining, Metallurgy and Petroleum Canadian Institute of Mining, Metallurgy and Petroleum – Toronto Branch Institute of Materials, Minerals and Mining, Affiliate Member

4. I have worked as a consultant in the minerals industry for 32 years, principally in the areas of mineral commodity market analysis and management of consulting assignments. My experience includes analysis of markets for base and precious minerals, industrial minerals, coal and uranium and management of the preparation of independent technical and due diligence reports.

5. I am familiar with NI 43-101 and, by reason of education, experience and professional registration, I fulfill the requirements of a Qualified Person as defined in NI 43-101;

6. I have participated in the compilation of the report and take responsibility for Section 18.11 of the report entitled “Independent Technical Report on The Results of a Preliminary Economic Assessment, Gurvabulag Uranium Deposit, Mongolia”;

7. Previously, I participated in the compilation of an internal economic assessment of the Gurvanbulag property on behalf of Western Prospector, dated September 9, 2005;

8. I am independent of the issuer under Section 1.4 of NI 43-101;

9. I have read NI 43-101 and the portion of this report for which I am responsible have been prepared in compliance with that instrument;

10. As of the date of this certificate, to the best of my knowledge, information and belief, the technical report contains all scientific and technical information that is required to be disclosed to make the technical report not misleading.

Dated this 27th day of November, 2007

“Jane Spooner”

Jane Spooner, M.Sc., P.Geo. Principal Micon International Limited

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Certificate for Mani Verma

I, Mani Verma do hereby certify that:

1. I am an associate of Micon International Limited, Suite 900, 390 Bay Street, Toronto, Ontario M5H 2Y2, tel. (416) 362-5135, fax (416) 362-5763, e-mail [email protected];

2. I hold the following academic qualifications:

B. Eng. (Mining) Sheffield University, UK 1963 M. Eng. (Mineral Economics) McGill University 1981

3. I am a registered Professional Engineer with the Professional Engineers of Ontario, registration no. 48070015; as well, I am a member in good standing of other technical associations and societies, including:

The Canadian Institute of Mining, Metallurgy and Petroleum

4. I have worked as a mining engineer in the minerals industry for over 25 years as a Consultant and having held senior management roles from Executive Vice President to Chief Operating Officer of a public mining company, 1996-2002;

5. I have read NI 43-101 and Form 43-101F1 and, by reason of education, experience and professional registration, I fulfill the requirements of a Qualified Person as defined in NI 43-101;

6. I am the author of, and am responsible for, Sections 18.1 through 18.5 and Sections 18.12 through 18.15 of the technical report entitled “Independent Technical Report on The Results of a Preliminary Economic Assessment, Gurvabulag Uranium Deposit, Mongolia”;

7. Previously, I participated in the preparation of an internal economic assessment of the Gurvanbulag property on behalf of Western Prospector, dated September 9, 2005;

8. I am independent of the issuers under Section 1.4 of NI 43-101;

9. I have read NI 43-101 and the portions of this report for which I am responsible have been prepared in compliance with that instrument;

Dated this 27th day of November, 2007

“Mani M. Verma”

Mani M. Verma, P.Eng.

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