View the Lake Maitland National Instrument 43-101 - Mega Uranium

Mega Uranium Ltd Lake Maitland National Instrument 43-101 Technical Report Report Prepared for

Mega Uranium Ltd

Frontispiece: Location of Mega Uranium’s Lake Maitland Uranium Deposit

Prepared by

MEG003 September 2009

SRK Consulting (Australasia) Pty Ltd Reg’d No ABN 56 074 271 720 Trading as SRK Consulting

SRK Consulting Project: MEG003 – Lake Maitland NI43-101 Report

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Lake Maitland National Instrument 43-101 Technical Report

Mega Uranium Ltd 57 Havelock Street West Perth WA 6005

SRK Consulting (Australasia) Pty Ltd 10 Richardson Street West Perth WA 6005

Daniel Guibal Corporate Consultant – SRK Consulting

MEG003 September 2009 Peter Gleeson Principal Consultant –SRK Consulting

Matthew Wheeler Geology Manager – Mega Uranium Ltd

Daryl Evans Principal Metallurgist – IMO Pty Ltd HERO/GLEE/GUIB/WILL/mool

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Table of Contents 3

Summary ....................................................................................................... 1

4

Introduction .................................................................................................. 3 4.1 4.2

Terms of reference ..........................................................................................3 Information sources .........................................................................................3

5

Reliance on Other Experts ........................................................................... 5

6

Property Description and Location ............................................................. 6 6.1

Property details ................................................................................................6 6.1.1

6.2 6.3

Tenement schedule ...................................................................................... 8

Agreements and royalties ..............................................................................15 Permits and obligations .................................................................................17 6.3.1 6.3.2 6.3.3

Minimum expenditure commitments ........................................................... 17 Bonds.......................................................................................................... 18 Environmental liabilities .............................................................................. 18

7

Accessibility, Climate, Local Resources, Infrastructure and Physiography .............................................................................................. 19

8

History ......................................................................................................... 20

9

Geological Setting ...................................................................................... 21

10

Deposit Type ............................................................................................... 22

11

Mineralisation ............................................................................................. 23

12

Exploration .................................................................................................. 24

13

Drilling ......................................................................................................... 25 13.1 13.2

Redport Ltd 2005 aircore drilling....................................................................25 Mega Uranium Ltd 2007/2008 drilling ............................................................26 13.2.1 13.2.2 13.2.3

13.3 13.4

14

Drillhole collar surveying ................................................................................31 Results ...........................................................................................................32

Sampling Method and Approach............................................................... 35 14.1

Downhole gamma logging .............................................................................35 14.1.1 14.1.2 14.1.3

14.2 14.3

Logging methods ........................................................................................ 35 Calibrations and verifications...................................................................... 35 Calculation of eU3O8 ................................................................................... 36

Geological logging .........................................................................................37 14.2.1 14.2.2

Redport Ltd ................................................................................................. 37 Mega Uranium Ltd ...................................................................................... 37

Sampling ........................................................................................................37 14.3.1 14.3.2

15

Aircore drilling ............................................................................................. 26 Sonic core drilling ....................................................................................... 30 Safety .......................................................................................................... 31

Redport Ltd ................................................................................................. 37 Mega Uranium Ltd ...................................................................................... 38

Sample Preparation, Analysis and Security ................................................ 39 15.1

Sampling ........................................................................................................39 15.1.1

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Redport Ltd ................................................................................................. 39

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15.1.2 15.1.3 15.1.4

16

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Mega Uranium Ltd ...................................................................................... 39 Adequacy of analytical procedures............................................................. 42 Bulk density determinations........................................................................ 42

Data Verification ......................................................................................... 43 16.1

Review of downhole gamma vs chemical assays ..........................................43 16.1.1 16.1.2 16.1.3 16.1.4 16.1.5

16.2

QA/QC ...........................................................................................................55 16.2.1 16.2.2

16.3 16.4

Basic statistical analysis ............................................................................. 44 Regression analysis ................................................................................... 48 QQ plot analysis ......................................................................................... 50 Downhole comparison ................................................................................ 52 Conclusions of comparative analysis ......................................................... 55 Assessment of drillhole LMAC0012............................................................ 56 Repeat logging............................................................................................ 57

Twinned holes................................................................................................59 QA/QC of chemical assays ............................................................................60 16.4.1 16.4.2 16.4.3

Blanks ......................................................................................................... 60 Standards ................................................................................................... 60 Pulp duplicates ........................................................................................... 60

17

Adjacent Properties.................................................................................... 61

18

Mineral Processing and Metallurgical Testing ......................................... 63 18.1 18.2

Introduction ....................................................................................................63 Sighter testwork .............................................................................................63 18.2.1 18.2.2 18.2.3 18.2.4 18.2.5 18.2.6 18.2.7 18.2.8 18.2.9

18.3

Scoping testwork ...........................................................................................80 18.3.1 18.3.2 18.3.3 18.3.4 18.3.5 18.3.6 18.3.7 18.3.8 18.3.9 18.3.10 18.3.11

18.4

Introduction ................................................................................................. 80 Scoping testwork samples .......................................................................... 80 Comminution testing ................................................................................... 80 Beneficiation testing ................................................................................... 81 Reverse sulphate flotation .......................................................................... 84 Ambient temperature pre-leaching ............................................................. 84 Reverse flotation and elevated temperature alkaline leaching ................... 85 Heap leach.................................................................................................. 85 Slurry rheology............................................................................................ 86 Uranium refining and recovery ................................................................... 88 Flowsheet testing ........................................................................................ 89

Process flowsheet..........................................................................................89 18.4.1 18.4.2 18.4.3 18.4.4

19

Sighter testwork samples ........................................................................... 63 Chemical characterisation .......................................................................... 66 Physical characterisation ............................................................................ 66 Mineralogy .................................................................................................. 67 Uranium alkaline leach extraction............................................................... 68 Alkaline leach carbonate consumption ....................................................... 76 Chloride content.......................................................................................... 76 Conclusions ................................................................................................ 77 Recommendations ...................................................................................... 79

Process flow ............................................................................................... 89 Process description .................................................................................... 90 Conclusions ................................................................................................ 91 Recommendations ...................................................................................... 92

Mineral Resource and Mineral Reserve Estimates .................................. 94 19.1

Geological modelling and domaining .............................................................94 19.1.1 19.1.2

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Geology modelling ...................................................................................... 94 Grade domaining ........................................................................................ 95

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19.2

Estimation ......................................................................................................97 19.2.1 19.2.2 19.2.3 19.2.4 19.2.5 19.2.6 19.2.7 19.2.8

19.3 19.4 19.5 19.6 19.7

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Database .................................................................................................... 97 Densities ..................................................................................................... 98 Disequilibrium ........................................................................................... 101 Statistical analysis of resource domains .................................................. 102 Application of upper cuts .......................................................................... 109 Variography .............................................................................................. 109 Block modelling......................................................................................... 114 Estimation/Kriging ..................................................................................... 115

Classification................................................................................................118 Resource summary......................................................................................119 Model limitations ..........................................................................................121 Recommendations .......................................................................................121 Modifying factors..........................................................................................122

20

Other Relevant Data and Information ..................................................... 123

21

Interpretation and Conclusions .............................................................. 124

22

Recommendations ................................................................................... 125

23

References ................................................................................................ 126

24

Data and Signature Pages ....................................................................... 127

25

Additional Requirements for Technical Reports on Development Properties and Production Properties .................................................... 131

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List of Tables Table 3-1: Table 3-2: Table 4-1: Table 6-1: Table 6-2: Table 13-1: Table 13-2: Table 13-3: Table 13-4: Table 15-1: Table 15-2: Table 15-3: Table 15-4: Table 15-5: Table 15-6: Table 16-1: Table 16-2: Table 16-3: Table 16-4: Table 17-1: Table 18-1: Table 18-2: Table 18-3: Table 18-4: Table 18-5: Table 18-6: Table 18-7: Table 18-8: Table 18-9: Table 18-10: Table 18-11: Table 18-12: Table 18-13: Table 18-14: Table 18-15: Table 18-16: Table 19-1: Table 19-2: Table 19-3: Table 19-4: Table 19-5: Table 19-6: Table 19-7: Table 19-8 Table 19-9:

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Indicated Resource .......................................................................................2 Inferred Resource .........................................................................................2 Qualified Persons..........................................................................................3 Tenement schedule ....................................................................................10 Summary of the LMUP mining tenure conditions ........................................14 Summary statistics for resource infill aircore and sonic drilling programmes completed by Mega during late 2007 and 2008 .....................28 Drillhole collar survey pick-ups 2008 ..........................................................32 Lake Maitland survey control points ............................................................32 Summary of significant intercepts from Mega aircore drilling 2007/08 .......33 Metallurgical sighter test samples – Ultratrace multi-element analysis details..........................................................................................................40 Metallurgical sighter test samples – Genalysis multi-element analysis details..........................................................................................................40 Mega aircore - Actlabs analytical details .....................................................41 Mega aircore - Genalysis analytical details .................................................41 Certified standards used by Mega ..............................................................41 Certified reference material used by Actlabs Pacific ...................................42 Mega 0.5 m composites summary statistics ...............................................44 Redport 0.5 m composites summary statistics ...........................................44 Statistical summary of assays and each log for LMAC0012 .......................56 Statistical summary of assays and repeats for all holes .............................57 Summary of adjacent resources .................................................................62 Sample matrix used to select sighter testwork samples .............................64 Sighter test sample locations ......................................................................64 Significant analytes sighter sample ICP head analysis...............................66 Physical characterisation by location ..........................................................67 Modal mineralogy........................................................................................67 LMAC0134 leach kinetic data .....................................................................72 LMAC0307 and LMAC0162 leach kinetic data ...........................................73 LMAC0314 leach kinetic data .....................................................................74 LMAC0336 and LMAC0567 leach kinetic data ...........................................75 Estimated leach sodium carbonate consumption rates ..............................76 Soluble chloride in leach solutions ..............................................................77 Preliminary comminution indices competent calcrete .................................81 Size x assay metal deportment ...................................................................82 Colorimetric beneficiation visual markers ...................................................83 LMAC0314 leach performance and reverse sulphate flotation ...................85 Heap leach performance LMAC0314 +355µm ...........................................86 Lithofacies estimated into S-Grid ................................................................94 Sonic core density summary statistics by lithology ...................................100 Density summary statistics by lithological grouping ..................................100 Density used by rock type .........................................................................101 Statistics for 0.25 m composites by geological domain ............................103 Variogram parameters ..............................................................................113 Indicated Resource ...................................................................................119 Inferred Resource .....................................................................................119 Indicated Resource at 100 ppm eU3O8 cut-off – by lithology ....................120

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Table 19-10: Indicated Resource at 200 ppm eU3O8 cut-off – by lithology ....................120 Table 19-11: Inferred Resource at 100 ppm eU3O8 cut-off – by lithology .......................120 Table 19-12: Inferred Resource at 200 ppm eU3O8 cut-off – by lithology .......................121

List of Figures Figure 6-1: Figure 6-2: Figure 6-3: Figure 13-1: Figure 13-2: Figure 13-3: Figure 13-4: Figure 13-5: Figure 16-1: Figure 16-2: Figure 16-3: Figure 16-4: Figure 16-5: Figure 16-6: Figure 16-7: Figure 16-8: Figure 16-9: Figure 16-10: Figure 16-11: Figure 17-1: Figure 18-1: Figure 18-2: Figure 18-3: Figure 18-4: Figure 18-5: Figure 18-6: Figure 18-7: Figure 18-8: Figure 18-9: Figure 18-10 Figure 18-11:

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Summary location map of LMUP detailed in Table 6-1.................................7 Location LMUP resource outline and tenements ..........................................8 Map of LMUP tenements ..............................................................................9 Wallis Drilling’s Mantis 75 track mounted drilling rig used to carry out resource drilling by Redport Ltd in 2005 .....................................................26 National Drilling’s Morooka track mounted KL-150 aircore rig used to carry out resource drilling............................................................................28 Drillhole location plan of Mega 2007/08 and Redport 2005 drilling with current resource limits.................................................................................29 Boart Longyear’s Minisonic track mounted sonic core drill rig used by Mega in 2008 ..............................................................................................31 Downhole Gamma grade thickness image of Mega 2007/08 aircore drilling and Redport 2005 drilling ...............................................................34 Histograms comparing chemical and gamma composited data sets for the Mega drilling..........................................................................................46 Histograms comparing chemical and gamma composited data sets for the Redport drilling ......................................................................................47 Scatter plot for Mega 0.5 m samples – downhole chemical vs gamma samples.......................................................................................................49 Scatter plot for Redport 0.5 m samples – downhole chemical vs gamma samples..........................................................................................50 QQ plot for gamma vs chemical assays – Mega data set...........................51 QQ plot for gamma vs chemical assays – Redport data set .......................52 Downhole graphs for Mega holes showing gamma assay values vs chemical assay values ................................................................................53 Downhole graphs for Redport holes showing gamma assay values vs chemical assay values ................................................................................54 Line chart of radiometric and assay data for LMAC0012 ............................57 Scatter plot of radiometric repeat data – all holes.......................................58 Twinned holes LMAC0135/LMAC1066 .......................................................59 Map of the region around Lake Maitland showing significant nearby uranium deposits.........................................................................................62 Sighter test sample locations ......................................................................65 Dolomite, clay and sulphate distribution .....................................................68 Final metal extraction and head grade........................................................69 Final metal extraction and head grade ratio................................................69 Uranium extraction and strontium head grade ............................................70 Uranium extraction and impact of sulphate .................................................70 Uranium extraction and impact of celestine and gypsum ...........................71 Alkaline leach kinetic data LMAC0134........................................................72 Alkaline leach kinetic data LMAC0307........................................................73 Alkaline leach kinetic data LMAC0314........................................................74 Alkaline leach kinetic data LMAC0336 and LMAC0567..............................75

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Figure 18-12: Figure 18-13: Figure 18-14: Figure 18-15: Figure 18-16: Figure 18-17: Figure 18-18: Figure 19-1: Figure 19-2: Figure 19-3: Figure 19-4: Figure 19-5: Figure 19-6: Figure 19-7: Figure 19-8: Figure 19-9: Figure 19-10: Figure 19-11: Figure 19-12: Figure 19-13: Figure 19-14: Figure 19-15: Figure 19-16: Figure 19-17:

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Alkaline leach sodium carbonate consumption rate ....................................76 Scrubbing and attritioning metal recovery and yield ...................................83 Heap leach performance LMAC0314 +355µm ...........................................86 Slurry rheology fresh feed -300µm..............................................................87 Slurry rheology alkaline leach feed .............................................................87 Slurry rheology alkaline leach tailings .........................................................88 Proposed Lake Maitland block flow diagram ..............................................89 Plan and 3D view of the base of channel surface constructed in Geomodeller ...............................................................................................95 3D view looking NE along the channel showing the categorical modelling of rock types in the channel ........................................................96 3D view of the combined GoCAD categorical geology model constrained by LeapfrogTM generated 100 ppm grade shell .......................97 CEC trench locations sampled by Mega for calcrete density testwork .......99 Scatterplot showing gamma eU3O8 values plotted against DFN U3O8 values........................................................................................................102 Frequency histogram and log probability plot for all domains in the general Channel domain ..........................................................................104 Frequency histogram and log probability plot for Calcrete domain ...........105 Frequency histogram and log probability plot for Calcrete Clay domain ...106 Frequency histogram and log probability plot for Clay domain .................107 Frequency histogram and log probability plot for Sandy Clay/Clayey Sand domain .............................................................................................108 Directional variography (actual and fitted theoretical) for the Global Channel domain ........................................................................................110 Directional variography (actual and fitted theoretical) for the Calcrete domain ......................................................................................................110 Directional variography (actual and fitted theoretical) for the Clay domain ......................................................................................................111 Directional variography (actual and fitted theoretical) for the Sandy Clay/Clayey Sand domain.........................................................................111 Plan view of resource block model grid with drillhole collar locations overlain .....................................................................................................114 Plan and 3D view of the S-Grid block model domained by geology .........115 Plan and 3D view of the Kriged estimates for eU3O8 on the channel block model ...............................................................................................117

List of Appendices Appendix 1: Appendix 2: Appendix 3:

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Scatter plots of repeat downhole radiometric measurements Twinned drillholes downhole graphs Apparent SG determinations on dry calcrete rock specimens

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Summary

This report details the updated mineral resource for the Lake Maitland Uranium Project („LMUP‟). Mega Uranium Ltd („Mega‟) is required under National Instrument 43-101 – Standards for Mineral Projects of the Canadian securities regulatory authorities or regulators – to file a technical report following their announcement of a resource update for the LMUP on 8 July 2009. This report, written collaboratively between Mega and SRK Consulting Australasia Pty Ltd („SRK‟), fulfils that obligation. Mega owns 100% of the LMUP and operates the project as Mega Redport Pty Ltd and its wholly owned subsidiary, Redport Exploration Pty Ltd. Mega acquired the project through their acquisition of Redport Ltd („Redport‟) in 2006. The LMUP is located in central Western Australia at latitude 27 10‟ 9‟‟ S, longitude 121 05‟ 46‟‟ E, approximately 740 km northeast of Perth, the capital city of Western Australia. The current project area comprises 13 exploration licences, 20 prospecting licences and 4 granted mining leases covering a total area of approximately 933 km2. The Lake Maitland deposit was first identified in a regional radiometric survey in 1967. Between discovery and the early 1980s, five companies were active in evaluating the project. After this time all work ceased until Acclaim Uranium took up the project in the late 1990s. Redport began drilling in 2005 and Mega have continued to evaluate the project since 2006. The Lake Maitland deposit lies within the Yandal Greenstone Belt of the Archean Yilgarn Craton. The deposit comprises a series of sediment and evaporite layers formed within a playa lake. Typical stratigraphy grades from basal red-brown silts and sands into calcrete which is overlain by further clays, silts and sands and topped by a gypsiferous unit. Locally the sedimentary facies are variable and average total thickness is in the order of 10 m. Uranium mineralisation, in the form of carnotite, is associated with calcrete, clay and sandy clay units. The local region of Western Australia appears well endowed with calcrete-hosted uranium systems as it has favourable source lithologies and a suitable climate. A revised resource estimate has been undertaken subsequent to extensive resource evaluation drill programs being completed by Mega in late 2007 and 2008. The updated resource relies solely on drilling data of Mega and Redport. The current resource drillhole database comprises 1,441 holes for a total of 17,451 m, the majority of which are aircore. Drillhole coverage at 100 mN by 100 mE spacing and 200 mN x 100 mE spacing has been achieved for the majority of the resource area. All holes have been geophysically logged using calibrated total gamma probes and the results converted to equivalent U3O8 (eU3O8) grades. Although only eU3O8 data were used in the estimate, over 2,715 chemical assays have been completed. Analysis of the assay results shows a slight negative bias of the gamma results and hence the grade estimate (eU3O8) is considered slightly conservative. Disequilibrium studies completed to date support the use of the total gamma technique as the main determinate of the U3O8 grades. In support of the updated resource, some sighter metallurgical work, followed by scoping level work, has been performed. In December 2008, 27 samples from aircore drilling were submitted for sighter work aiming to assess suitability of carbonate leaching for treatment of the resource. Results proved insightful and allowed for further scoping level tests. These tests were performed in April 2009. Using a two tonne bulk sample, the tests were targeting a base case process flowsheet which can be used to design future testing and in performing an economic analysis of the project.

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The main difference from previous resource models is the application of a 3D geological model to constrain the eU3O8 estimate. By using an inverse distance algorithm, the different lithofacies within the deposit were effectively modelled in 3D. The geological domains were then used to domain the eU3O8 for estimation and were assigned appropriate densities. The principal reason for the improved resource classification is higher confidence in density derived from additional test work, better modelling of the rock types and elimination of historical data. Grade estimation of the resource was performed separately for each geological domain using Ordinary Kriging of the downhole radiometric eU3O8 data. The new resource is tabulated in Table 3-1 and Table 3-2. When the cut-off grade is lifted from 100 ppm eU3O8 to 200 ppm eU3O8, a significant decrease in the Indicated Resource tonnage of 9.85 Mt (34%) and a significant increase in the Indicated Resource grade from 376 ppm eU3O8 to 497 ppm eU3O8 (32%) corresponds with a much smaller decrease in contained metal of 3.1 Mlb (13%). Table 3-1: Indicated Resource Cut-off grade (ppm eU3O8)

Ore tonnage (kt)

Average grade (ppm eU3O8)

Contained eU3O8 (metal tonnes)

Contained eU3O8 (lbs metal x 106)

100

28,751

376

10,810

23.83

150

23,445

426

9,987

22.01

200

18,901

497

9,394

20.71

250

14,976

569

8,521

18.78

500

6,077

882

5,360

11.81

Table 3-2: Inferred Resource Cut-off grade (ppm U3O8)

Ore tonnage (kt)

Average grade (ppm eU3O8)

Contained U3O8 (metal tonnes)

Contained U3O8 (lbs metal x 106)

100

3,574

274

979

2.16

150

2,807

312

876

1.93

200

1,922

374

719

1.58

250

1,397

433

605

1.33

500

337

759

256

0.61

Given the significant upgrade in resource classification from Inferred to Indicated category, the authors recommend that the LMUP proceed to Feasibility Study at the earliest opportunity. This recommendation is in line with Mega‟s stated objective of achieving production at Lake Maitland in Q4, 2011.

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Introduction

At the request of Mr Matthew Wheeler, Manager Geology (Lake Maitland Uranium Project) of Mega Uranium Ltd („Mega‟), SRK Consulting Australasia Pty Ltd („SRK‟) was commissioned in July 2009 to prepare a Technical Report on an update of the mineral resource for the Lake Maitland Uranium Project („LMUP‟) compliant with the requirements of the National Instrument 43-101. Under Canada‟s National Instrument 43-101 (NI43-101) Standards of Disclosure for Mineral Projects, Mega is required to file a Technical Report following their announcement of a resource increase for the LMUP on 8 July 2009. This report aims to discharge Mega‟s obligation as described in Section 4.2 (1) (j) (ii) of NI43-101. This report follows the outline presented in Form 43-101F1 describing technical reports. Unless otherwise specified, the monetary unit used in this Report is Australian Dollars.

4.1

Terms of reference

SRK has been commissioned by Mega to assist in compilation of this Technical Report for the LMUP and where necessary, to act as Independent Qualified Persons. Given that this is not a first time disclosure nor represents a 100% change or greater from the previous resource estimate, there is no specific requirement for submission by Independent Qualified Persons as described in Section 5.3 of NI43-101. In March 2009, SRK was employed by Mega to update the resource estimate of the LMUP incorporating geological modelling and resource estimation. A report by SRK Principal Consultant Peter Gleeson describes the resource estimate in detail and is the basis for Mega‟s reported resource increase on 8 July 2009. The basis of this Technical Report is the updated resource estimate. Mega and SRK have worked collaboratively on this report with considerable input from the LMUP Manager Geology, Matthew Wheeler, and a number of Mega‟s other technical consultants described in Section 5. The Qualified Persons (QPs) for this Technical Report are listed in Table 4-1. Table 4-1: Qualified Persons Name

Affiliation

Discipline

Qualifications and registration

Site inspection

Daniel Guibal

SRK

Report compilation

MSc, MAusIMM (CP)

N/A

Peter Gleeson

SRK

Resource estimation

MSc, MAIG

March 2009 for 2 days

Matthew Wheeler

Mega

BSc (Hons), MAIG

Regular visits

Daryl Evans

IMO

Report compilation MetallurgyProcessing

BSc, MAusIMM

March 2009 for 1 day

4.2

Information sources

This Technical Report updates the previous NI43-101 technical report titled, „First Time Disclosure: Mega Uranium Ltd. Mineral Resource for Lake Maitland Uranium Deposit‟ prepared for Mega by Hellman & Schofield Pty Ltd in February 2007 (Hellman & Schofield, 2007). Where no material change has occurred, this report was used as a basis.

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The updated resource estimate is based on the SRK report titled, „Lake Maitland Resource Estimate‟ (Gleeson, 2009). Other information sources are referenced as appropriate in the text.

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Reliance on Other Experts

With respect to the tenements listed in Table 6-1, SRK has not conducted any separate study as to their standing, nor has SRK reviewed any legal agreements as to the rights of Mega to the tenements. Mr Marcus Walter of M&M Walter Consulting, Mega‟s tenement manager, was engaged to prepare a tenement report for Lake Maitland and the authors have relied on this report. Matters relating to legal agreements, royalties and Native Title have relied on information provided by Mr Richard Homsany, Solicitor - Mining & Energy, DLA Phillips Fox.

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6

Property Description and Location

6.1

Property details

The LMUP is centred at latitude 27 10‟ 9‟‟ S, longitude 121 05‟ 46‟‟ E, approximately 740 km northeast of Perth, the capital city of Western Australia (see Figure 6-1). The project area is serviced by the towns of Wiluna (108 km northwest) and Leinster (92 km south-southwest). The current LMUP granted mineral tenure comprises 13 exploration licences, 20 prospecting licences and 4 mining leases covering a total area of approximately 933 km2. The reported uranium resource (100 ppm eU3O8 cut-off) is covered by E53/947, E53/1099, E53/1256 and P53/1259-1261 inclusive (see Figure 6-2). A mining lease application (M53/1089) was lodged on 13 December 2008 to cover the majority of the currently identified uranium resource. The boundary of M53/1089 incorporates all the ground covered by E53/947 and P53/1257-1263 inclusive and the majority of the ground covered by E53/1099 and P53/1256. The majority of the LMUP tenements are located on the Barwidgee and Wonganoo pastoral properties. In addition to the mining lease application, a miscellaneous licence (L53/152) application has been lodged to facilitate exploration for groundwater. Exploration licence boundaries are defined by lines of predetermined latitude and longitude. These areas are known as graticules with each graticule being one minute of latitude by one minute of longitude. Each graticule is referred to as a „block‟ under the West Australian Mining Act 1978. Prospecting licences and mining leases are physically marked-out (pegged) on the ground. The position of all tenement boundaries are geodetically referenced to the Geodetic Datum of Australia 1994 (GDA94) with coordinates referenced to the Map Grid of Australia 1994 (MGA94).

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Figure 6-1: Summary location map of LMUP detailed in Table 6-1

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Figure 6-2: Location LMUP resource outline and tenements

6.1.1

Tenement schedule

Mega owns 100% of the LMUP through its acquisition of formerly Australian-listed company Redport Ltd (Redport). Mega operates the LMUP as Mega Redport Pty Ltd („Mega Redport‟) and its wholly owned subsidiary Redport Exploration Pty Ltd („RedEx‟). All granted licences and applications are registered in the name of RedEx, with the exception of granted exploration licences E53/1026, E53/1213, E53/1214, which are registered in the name of Yandal Metals Pty Ltd. A Schedule of Tenements has been prepared by Mega Redport and is presented in Table 6-1. The schedule has been cross-checked against the schedule provided by the tenement managers, M&M Walter Consulting of Perth.

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Figure 6-3: Map of LMUP tenements

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Table 6-1: Tenement schedule Lease

Registered holder

E 37/895 E 37/970 E 37/971 E 53/1026 E 53/1060 E 53/1099 E 53/1210 E 53/1211 E 53/1213 E 53/1214 E 53/1441 E 53/1442 E 53/947 L 53/152 M 53/1089 M 53/574 M 53/575 M 53/578 M 53/579 P 37/6943 P 53/1252 P 53/1253 P 53/1254 P 53/1255 P 53/1256 P 53/1257 P 53/1258 P 53/1259 P 53/1260 P 53/1261 P 53/1262 P 53/1263 P 53/1324 P 53/1336 P 53/1337 P 53/1338 P 53/1339 P 53/1340 P 53/1341

Redport Exploration Pty Ltd Redport Exploration Pty Ltd Redport Exploration Pty Ltd Yandal Metals Pty Ltd Redport Exploration Pty Ltd Redport Exploration Pty Ltd Redport Exploration Pty Ltd Redport Exploration Pty Ltd Yandal Metals Pty Ltd Yandal Metals Pty Ltd Redport Exploration Pty Ltd Redport Exploration Pty Ltd Redport Exploration Pty Ltd Redport Exploration Pty Ltd Redport Exploration Pty Ltd Redport Exploration Pty Ltd Redport Exploration Pty Ltd Redport Exploration Pty Ltd Redport Exploration Pty Ltd Redport Exploration Pty Ltd Redport Exploration Pty Ltd Redport Exploration Pty Ltd Redport Exploration Pty Ltd Redport Exploration Pty Ltd Redport Exploration Pty Ltd Redport Exploration Pty Ltd Redport Exploration Pty Ltd Redport Exploration Pty Ltd Redport Exploration Pty Ltd Redport Exploration Pty Ltd Redport Exploration Pty Ltd Redport Exploration Pty Ltd Redport Exploration Pty Ltd Redport Exploration Pty Ltd Redport Exploration Pty Ltd Redport Exploration Pty Ltd Redport Exploration Pty Ltd Redport Exploration Pty Ltd Redport Exploration Pty Ltd

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Legal units Block Block Block Block Block Block Block Block Block Block Block Block Block Hectare Hectare Hectare Hectare Hectare Hectare Hectare Hectare Hectare Hectare Hectare Hectare Hectare Hectare Hectare Hectare Hectare Hectare Hectare Hectare Hectare Hectare Hectare Hectare Hectare Hectare

Legal area 36 14 25 5 2 22 26 25 70 6 29 50 4 116289 7327 11.5717 14.1247 676.678 26.6281 197 197 198 178 200 186 189 183 152 123 199 199 199 72.8917 108.904 188.604 174.683 174.662 170.319 178.939 Total =

Area (ha) 10941.000 4260.970 7596.970 1527.890 610.364 5839.870 6748.370 5714.880 19142.360 1832.510 8840.470 15264.560 788.802 116290.280 7326.790 11.572 14.125 676.685 26.629 197.011 197.001 198.579 177.707 198.285 187.801 189.197 182.968 152.090 122.771 198.810 198.791 199.537 72.894 108.907 188.608 174.683 174.664 170.324 178.941 216924.67

Area (km2) 109.4 42.6 76.0 15.3 6.1 58.4 67.5 57.1 191.4 18.3 88.4 152.6 7.9 1162.9 73.3 0.1 0.1 6.8 0.3 2.0 2.0 2.0 1.8 2.0 1.9 1.9 1.8 1.5 1.2 2.0 2.0 2.0 0.7 1.1 1.9 1.7 1.7 1.7 1.8 2169.2

Status Granted Granted Granted Granted Granted Granted Granted Granted Granted Granted Granted Granted Granted Application Application Granted Granted Granted Granted Granted Granted Granted Granted Granted Granted Granted Granted Granted Granted Granted Granted Granted Granted Granted Granted Granted Granted Granted Granted

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Application date 23/01/2007 04/02/2008 04/02/2008 29/01/2002 04/09/2002 11/08/2003 10/02/2006 10/02/2006 15/02/2006 15/02/2006 10/11/2008 10/11/2008 08/09/2000 04/11/2008 13/12/2008 14/02/1997 14/02/1997 14/02/1997 14/02/1997 28/03/2006 28/03/2006 28/03/2006 28/03/2006 28/03/2006 28/03/2006 28/03/2006 29/03/2006 20/03/2006 29/03/2006 29/03/2006 29/03/2006 29/03/2006 22/01/2007 25/01/2007 25/01/2007 25/01/2007 21/01/2007 21/01/2007 25/01/2007

25/03/2008 30/12/2008 29/12/2008 27/07/2005 18/08/2005 22/06/2005 18/01/2007 10/01/2007 05/01/2007 05/01/2007 23/07/2009 23/07/2009 17/11/2004

Expiry date 24/03/2013 29/12/2013 28/12/2013 26/07/2010 17/08/2010 21/06/2010 17/01/2012 09/01/2012 04/01/2012 04/01/2012 22/07/2014 22/07/2014 16/11/2009

14/01/2009 14/01/2009 14/01/2009 27/11/2008 20/03/2007 31/01/2007 31/01/2007 31/01/2007 31/01/2007 31/01/2007 31/01/2007 31/01/2007 31/01/2007 31/01/2007 31/01/2007 31/01/2007 06/03/2007 29/07/2008 12/06/2008 12/06/2008 12/06/2008 12/06/2008 12/06/2008 12/06/2008

13/01/2030 13/01/2030 13/01/2030 26/11/2029 19/03/2011 30/01/2011 30/01/2011 30/01/2011 30/01/2011 30/01/2011 30/01/2011 30/01/2011 30/01/2011 30/01/2011 30/01/2011 30/01/2011 05/03/2011 28/07/2012 11/06/2012 11/06/2012 11/06/2012 11/06/2012 11/06/2012 11/06/2012

Grant date

Expenditure commitment $36,000.00 $20,000.00 $25,000.00 $20,000.00 $20,000.00 $33,000.00 $26,000.00 $25,000.00 $70,000.00 $20,000.00 $29,000.00 $50,000.00 $20,000.00 $0.00 $0.00 $10,000.00 $10,000.00 $67,700.00 $10,000.00 $7,880.00 $7,880.00 $7,920.00 $7,120.00 $8,000.00 $7,440.00 $7,560.00 $7,320.00 $6,080.00 $4,920.00 $7,960.00 $7,960.00 $7,960.00 $2,920.00 $4,360.00 $7,560.00 $7,000.00 $7,000.00 $6,840.00 $7,160.00 $630,540.00

DMP rental $4,268.88 $1,660.12 $2,964.50 $922.90 $369.16 $4,060.76 $4,799.08 $4,614.50 $12,920.60 $1,107.48 $3,438.82 $5,929.00 $1,001.88 $0.00 $0.00 $187.44 $234.30 $10,574.74 $421.74 $455.07 $455.07 $457.38 $411.18 $462.00 $429.66 $436.59 $422.73 $351.12 $284.13 $459.69 $459.69 $459.69 $168.63 $251.79 $436.59 $404.25 $404.25 $395.01 $413.49 $67,493.91

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Mineral tenure Section 9 of the Mining Act 1978 states, “Except in the case of land alienated in fee simple before the 1st January, 1899 (in which case minerals other than gold, silver and precious metals are the property of the owner), all minerals are the property of the Crown.” Where the minerals are the property of the Crown, a mining title must be obtained from the Department of Mines and Petroleum (DMP) before any mining operations may be undertaken. All LMUP tenements are located on Crown land. For the purposes of the Mining Act 1978 the State of Western Australia is divided into various mineral fields, some further divided into districts. The LMUP tenements belong to the East Murchison (53) and Mt Margaret (37) mineral fields. The mining tenements available under the Mining Act 1978 are as follows: Prospecting Licences (Sections 40-56) Special Prospecting Licences for Gold (Sections 56A, 70 and 85B) Exploration Licences (Sections 57-69E) Retention Licences (Sections 70A-70M) Mining Leases (Sections 71-85) General Purpose Leases (Sections 86-90) Miscellaneous Licences (Sections 91-94). The basic provisions of each type of mining tenure included as part of the LMUP are summarised in Table 6-2 and further described below. Prospecting Licences The maximum area for a prospecting licence is 200 hectares. Prospecting licences must be marked out. Application is made to the Mining Registrar of the relevant Mineral Field. An application fee and rental is payable. There is no limit to the number of licences a person or company may hold, but a security (or bond) is required in respect of each licence. The term of a prospecting licence is 4 years, with the provision to extend for one further 4 year period. The holder of a prospecting licence may, in accordance with the licence conditions, extract or disturb up to 500 tonnes of material from the ground, including overburden, and the Minister may approve extraction of larger tonnages.

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Exploration Licences On 28 June 1991, a graticular boundary (or block) system was introduced for exploration licences. The minimum size of an exploration licence is one block, and the maximum size is 70 blocks, except in areas not designated as mineralised areas, where the maximum size is 200 blocks. An exploration licence is not marked out. Application is made to the Mining Registrar of the relevant Mineral Field. An application fee and rental is payable. There is no limit to the number of licences a person or company may hold but a security (or bond) is required in respect of each licence. The term of an exploration licence is 5 years. The Minister may extend the term in prescribed circumstances. At the end of both the third and fourth year of its term, the licensee is required to surrender 50% of the licence. For a licence applied for and granted after 10 February 2006, the surrender requirement is 40% at the end of the fifth year. The holder of an exploration licence may, in accordance with the licence conditions, extract or disturb up to 1,000 tonnes of material from the ground, including overburden, and the Minister may approve extraction of larger tonnages. Mining Lease The maximum area for a mining lease applied for before 10 February 2006 is 1,000 hectares. After this date, the size applied for is to relate to an identified orebody as well as an area for infrastructure requirements. Mining leases must be marked out. Application is made to the Mining Registrar of the relevant Mineral Field. An application fee and rental is payable. Pursuant to Section 74(1)(ca), an application for a mining lease shall be accompanied by a mining proposal OR a statement in accordance with Subsection (1a) and a mineralisation report that has been prepared by a Qualified Person. The statement under Subsection (1a) shall set out information regarding the mining operation likely to be carried out including: -

-

When mining is likely to commence The most likely method of mining The location and area of land that is likely to be required for the operation of the plant, machinery and equipment and for the other activities associated with those mining operations There is no limit to the number of mining leases a person or company may hold. The term of a mining lease is 21 years and may be renewed for further terms.

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The lessee of a mining lease may work and mine the land, take and remove minerals and undertake all things necessary to effectually carry out mining operations in, on or under the land, subject to conditions of title.

Miscellaneous Licence There is no maximum area for a miscellaneous licence. Miscellaneous licences must be marked out. A miscellaneous licence is for purposes such as a road, pipeline, water, as prescribed in the Regulations, or such other purposes as the Director General of the Department of Industry and Resources may approve. Application is made to the Mining Registrar of the relevant Mineral Field. An application fee and rental is payable. There is no limit to the number of licences a person or company may hold. The term of a miscellaneous licence is 21 years, and may be renewed for further terms. A miscellaneous licence can be applied for over, and can co-exist with, other mining titles.

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Table 6-2: Summary of the LMUP mining tenure conditions Fees Type

Prospecting Licence

Exploration Licence (Graticular)

Maximum area

200 ha

70 Blocks 200 Blocks (outside known mineralised areas)

Terms

4 years Renewable for 1 period of 4 years (for licences applied for after 10 February 2006)

5 years May extend for 2 periods of up to 2 years and further periods of 1 year for licences applied for prior to 10 February 2006. On or after this date term is 5 years, may extend for one period of 5 years and by a further period or periods of 2 years

Application

Rent on application

Rent including GST

$250.00

$2.10 per ha or part thereof Min $21.00

$2.31 per ha or part thereof Min $23.10

Years 1–3 $118.58 per block ($285.67 if for only 1 block) $1,115 ($260 if for 1 block only)

$107.80 per block ($259.70 if for only 1 block)

Years 4 and 5 $184.58 Years 6 and 7 $250.47 Year 8 on $474.32

Minimum annual expenditure

$40.00 per ha Min $2,000

Years 1–3 $1,000 per block • Min $10,000 for 1 block • Min $15,000 for 2–5 blocks • Min $20,000 for 6 to 20 blocks Escalating expenditure for Year 4 and thereafter $100 per ha

Mining Lease

N/A

21 years renewable

$375.00

$14.20 per ha or part thereof

$15.62 per ha or part thereof

Miscellaneous Licence (search for groundwater)

N/A

21 years renewable

$375.00

40c per ha or part thereof

44c per ha or part thereof

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Min $5,000 if 5 ha or less Otherwise $10,000

Covenant in lieu

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Agreements and royalties

Farm-in and joint venture On 18 June 2009, RedEx entered into agreements with the Japan Australia Uranium Resources Development Co Ltd (through its subsidiary JAURD International Lake Maitland Project Pty Ltd) (JAURD) and ITOCHU Corporation (through its subsidiary ITOCHU Minerals & Energy of Australia Pty Ltd, or IMEA) for a farm-in and joint venture in respect of RedEx's LMUP [which comprises tenements E53/947, E53/1099 and P53/1256-1263 (inclusive) and application M53/1089]. JAURD is a Japanese company mandated to acquire uranium resources in Australia on behalf of its shareholders, being three Japanese utilities – Kansai Electric Power Company Incorporated (50%), Kyushu Electric Power Company Incorporated (25%) and Shikoku Electric Power Company Incorporated (15%) – and ITOCHU Corporation (10%), which is one of the world's largest uranium trading houses. Under the farm-in agreement, JAURD and IMEA have agreed to contribute up to an aggregate amount of US$49,000,000 in three stages, in return for which JAURD and IMEA will be entitled to earn a 30% and a 5% interest in the LMUP, respectively. Initial payments by JAURD and IMEA will be used to fund exploration and feasibility studies in respect of the LMUP. Payments may be made by JAURD and IMEA at their discretion, and JAURD and IMEA are entitled to withdraw from the farm-in agreement at any stage prior to payment of the final instalments. In the event that IMEA elects to withdraw from the farm-in agreement, JAURD is entitled to make any outstanding contributions that would have otherwise been made by IMEA and to acquire IMEA's proposed 5% interest in the LMUP in addition to its own 30% interest. Upon payment of the final instalments by JAURD and IMEA (or JAURD alone) and their share of development costs incurred, the parties will form a joint venture to be known as the Lake Maitland Development Joint Venture, which will be governed by the joint venture agreement dated 18 June 2009 between RedEx, JAURD and IMEA. LMUP, a wholly owned subsidiary of Mega, will be the initial manager of the Lake Maitland Joint Venture, pursuant to a management agreement of the same date. Notification has been received from the Foreign Investment Review Board of Australia that there are no objections to the investment by JAURD and IMEA to acquire a 30% interest and a 5% interest in the LMUP respectively. The definitive agreements remain conditional upon the receipt of executed deeds with the royalty holders over certain tenements within the LMUP (see Royalty Agreements below). The parties have also entered into an option agreement with Mega Redport pursuant to which JAURD and IMEA have acquired an option to gain an interest in the other tenements within the LMUP [tenements E53/1060, E53/1210-1211, P37/6943, P53/1252-1255 (inclusive), P53/1324, P53/1336-1341 (inclusive), M53/574-575 and M53/578-579], and in Mega Redport's uranium rights under an agreement with Artemis Resources Ltd (Artemis), in return for the payment to RedEx of 30% and 5% respectively, of all expenditure incurred by RedEx from the date of the option agreement until the date of the exercise of the option in connection with the tenements that are the subject of the option agreement. In the event that IMEA elects not to exercise its option under the option agreement, JAURD is entitled to acquire both its own 30% interest and IMEA's 5% interest. Upon JAURD, or JAURD and IMEA, acquiring an interest in the tenements and the rights that are the subject of the option agreement, those tenements and rights will become assets of the Lake Maitland Joint Venture.

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JAURD and ITOCHU have lodged caveats over E53/947, E53/1060, E53/1099, E531210, E531211, P53/1252-P53/1263, P53/1324, P53/1336-1341, P37/6943, M53/574, M53/575, M53/578 and M53/579 which prevents RedEx from registering any dealings with respect to these licences. Royalty agreements Exploration licence 53/1060 was acquired by RedEx from Joydem Pty Ltd (Joydem) pursuant to a sale agreement dated 23 September 2005. Under that agreement, RedEx has agreed to pay to Joydem a gross royalty equal to 1% of the value of all minerals produced and sold from tenement E53/1060. Joydem has a registered caveat over tenement E53/1060 and is entitled, upon RedEx notifying it of its intention to establish a mining operation or commence a bankable feasibility study in respect of tenement E53/1060, to register a mortgage over that tenement to protect its royalty entitlement. Exploration licence 53/947 was acquired by RedEx from Coniston Pty Ltd (Coniston) pursuant to a sale agreement dated 23 September 2005. Under that agreement, RedEx has agreed to pay to Coniston a gross royalty equal to 1% of the value of all minerals produced and sold from tenement E53/947. Coniston has a registered caveat over tenement E53/947 and is entitled, upon RedEx notifying it of its intention to establish a mining operation or commence a bankable feasibility study in respect of tenement E53/947, to register a mortgage over that tenement to protect its royalty entitlement. Tenements E53/1210, E53/1211, P53/1252 - P53/1263 (inclusive), P53/1324, P53/1336 - P53/1341 (inclusive), P37/6943, M53/574, M53/575, M53/578 and M53/579 are the subject of a royalty deed between Newmont Yandal Options Pty Ltd (Newmont) and RedEx. Under that deed, RedEx agreed to pay to Newmont a 1% net smelter return royalty on all gold extracted from those tenements and 1% net smelter return royalty on any other metals extracted from those tenements. Newmont has the right to lodge a mortgage over those tenements which are the subject of the royalty deed. Newmont has assigned its entitlement to both the gold royalty and the other metals royalty to Franco-Nevada Australia Pty Ltd (Franco-Nevada) pursuant to a document entitled, 'Deed of Assignment, Assumption Maitland Royalty' dated 24 January 2008. Uranium rights Yandal Redport Resources Pty Ltd (now Yandal Metals Pty Ltd), a company which was formerly a wholly-owned subsidiary of Mega Redport (then Redport), was sold to Goldfields Consolidated Ltd (now Artemis) pursuant to a letter agreement dated 30 June 2006. Pursuant to that agreement, the parties agreed that Mega Redport would retain all rights in respect of uranium conferred under tenements E53/1026, E53/1213 and E53/1214, which are owned by Yandal Metals Pty Ltd, now a wholly-owned subsidiary of Artemis.

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Permits and obligations

In Australia, all minerals belong to the Crown. Under the Australian system of government the Commonwealth and State Governments are responsible for different aspects of the regulatory system. The Commonwealth Government is responsible for overall economic policy, tax, interest rates, foreign investment and corporate law, and for regulations regarding environmental and safety aspects of uranium mining and the sale of uranium product. The six States and the Northern Territory of Australia own and allocate mineral property rights for exploration and mining, regulate operations and collect royalties on minerals produced. The various regulatory authorities and other parties with responsibilities or interests in the area of the mining tenements are: Australian Government: Department of Environment, Water, Heritage and the Arts (DEWHA) Australian Government: Australian Safeguards and Non-Proliferation Office (ASNO) Australian Government: Department of Resources, Energy and Tourism (RET) Western Australian Environmental Protection Authority (EPA) Western Australian Department of Mines and Petroleum (DMP) Western Australian Radiological Council Western Australian Department of Environment and Conservation (DEC) Western Australian Department of Water (DoW) Shire of Wiluna Shire of Leonora Various Pastoral Lease Holders Central Desert Native Title Services Ltd. Before exploration can commence, an exploration or prospecting lease must be granted over the land to be explored. A Programme of Works (PoW) must be obtained from the Department of Mines and Petroleum (DMP) in Western Australia for the use of mechanised equipment on a mining lease. Mega has submitted a PoW to undertake a costeaning programme on E53/1099 and E53/947 as part of its ongoing feasibility studies and to conduct exploratory drilling on E53/1099, E53/947, P53/1256-1258, E53/1210 and E53/1211. A licence must be obtained from the Department of Water (DoW) prior to the construction of bores, or in order to take ground- or surface water. Mega has a number of licences in place to alter and construct bores over the LMUP.

6.3.1

Minimum expenditure commitments

Prospecting Licences, Exploration Licences and Mining Leases are subject to a prescribed minimum annual expenditure commitment enforced by the DMP. This requirement applies to granted tenements only and the labour cost of the tenement holders' own work on the tenement (contract equivalent) may be treated as expenditure. There is no prescribed annual expenditure for a Miscellaneous Licence; however, the Minister may determine the level of expenditure by condition on grant. If a licensee/lessee cannot fulfil the expenditure obligations, he/she may apply for exemption from all or part of the commitment. The combined annual expenditure for the LMUP tenements is $630,540.00. Annual rents to DMP total $67,493.91.

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Bonds

Unconditional Performance Bonds have been placed over some of the tenements of the LMUP. Details are outlined below: Date imposed

13-12-2005

Tenement

Details

E53/1099

The licensee arranging lodgement of an Unconditional Performance Bond executed by a Bank or other approved financial institution in favour of the Minister for State Development for due compliance with the environmental conditions of the lease in the sum of $20,000.

Bonds are intended to provide the State with a guaranteed access to funds so that necessary rehabilitation can be undertaken on mining tenements in cases where tenement holders fail to comply with environmental conditions placed on their tenements. The third party providing the Bond remains liable to the Minister even when the tenement holder(s) is/are in bankruptcy or liquidation. Cash bonds are not acceptable in lieu of Unconditional Performance Bonds. A Bond remains registered and enforceable until it is retired by the Minister, i.e. when he or she is satisfied that the relevant obligations have been met by the tenement holder(s). The Bond document is then returned to the relevant financial institution and written notification of the retirement is sent to the tenement holder(s) listed on the Bond document.

6.3.3

Environmental liabilities

The LMUP is subject to the Mining Act 1978 under which there are environmental liabilities. The use of mechanised equipment requires the approval of a PoW. There are a number of environmental conditions attached to the approval of a PoW. There are also a number of standard environmental conditions over each tenement.

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Accessibility, Climate, Local Resources, Infrastructure and Physiography

The LMUP is situated in the North Eastern Goldfields region of Western Australia. The project‟s location in a major mining centre provides good access by road, a nearby gas pipeline and ready source of experienced labour. The area is classed as semi-arid and is covered by low level scrub typical of the region. Lake Maitland is part of a low gradient drainage system incorporating temporal playa lakes. Further discussion of the accessibility, climate, local resources, infrastructure and physiography can be found in the previous Lake Maitland Technical Report (Hellman and Schofield, 2007) available on the SEDAR website (www.sedar.com).

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History

Seven companies have undertaken work at Lake Maitland (also known as Mt Joel) since the first radiometric survey by the Bureau of Mineral Resources in 1967. They are, in order of work: Australis Mining Asarco (Wiluna Gold Mines) Carpentaria Exploration Company (Mt Isa Mines) BP Minerals Australia Esso (Exxon Coal and Minerals) Acclaim Uranium Redport (now Mega Redport) Exploration by these companies comprised a range of techniques, including scintillometer traverses, auger drilling, RC drilling and trenching. Further discussion of the exploration history and previous resource estimates can be found in the previous Lake Maitland Technical Report (Hellman and Schofield, 2007) available on the SEDAR website (www.sedar.com). Details of Redport‟s and Mega‟s exploration are discussed in Section 12.

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Geological Setting

The LMUP lies within the Yandal Greenstone Belt of the Archean Yilgarn Craton. The uranium deposit is hosted within a series of sediment and evaporite layers formed within a playa lake. Typical stratigraphy grades from basal red-brown silts and sands into calcrete which is overlain by further clays, silts and sands and topped by a gypsiferous unit. Locally the sedimentary facies are variable and average total thickness is in the order of 10 m. Uranium mineralisation, in the form of carnotite, is associated with calcrete, clay and sandy clay units. Further discussion of the geological setting can be found in the previous Lake Maitland Technical Report (Hellman and Schofield, 2007) available on the SEDAR website (www.sedar.com).

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Deposit Type

Butt et al. (1984) classified the LMUP as a surficial calcrete-hosted uranium deposit. Calcretes (calcium and magnesium carbonates) form in arid to semi-arid terrains where evaporation promotes their deposition. In Western Australia the uranium and vanadium are sourced from Archean granitic and greenstone rocks, respectively, and precipitation of uranium minerals [typically carnotite, (K2(UO2)2(VO4)2.1-3H2O)] occurs as transporting fluids become depleted in carbonate and concentrated in vanadium through evaporation. Further discussion of the calcrete deposit type and their exploration can be found in the previous Lake Maitland Technical Report (Hellman and Schofield, 2007) available on the SEDAR website (www.sedar.com).

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Mineralisation

Uranium mineralisation at the LMUP is flat-lying and thin, averaging around 1.7 m in thickness, beneath approximately 1.5 - 2.0 m of sand, silt and other evaporites. The mineralisation has a large areal extent, approximately 5 km long (N-S) and around 2 km wide (E-W). The deposit is essentially crescent-shaped with three arms extending towards the west – the north-western, mid-western and south-western arms. The uranium mineralisation occurs as the mineral carnotite (K2(UO2)2(VO4)2.1-3H2O) (Section 18.2.4). The carnotite generally occurs within voids in the calcrete and as disseminations within the sands, silts and clays. Further description of the mineralisation can be found in the previous Lake Maitland Technical Report (Hellman and Schofield, 2007) available on the SEDAR website (www.sedar.com).

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Exploration

The exploration history of Lake Maitland is summarised in Section 8 (above), while drilling and associated studies carried out by Redport and then Mega are described in Sections 13-16 (below). Further detail of work undertaken by historical explorers is described in the previous Lake Maitland Technical Report (Hellman and Schofield, 2007) available on the SEDAR website (www.sedar.com).

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Drilling

The updated resource estimate described in this report relies on data derived from resource evaluation drilling undertaken by Redport in 2005 and Mega in 2007/08. The previous NI43-101 reported Inferred Resource estimate (Hellman and Schofield, 2007) relied upon drilling undertaken by Redport in 2005 and historical explorers between 1972 and 1979. As the historical drilling has not been relied upon for this revised resource estimate, it is not discussed further in this report. For full details of historical exploration and evaluation drilling, the reader should refer to the previous Lake Maitland Technical Report (Hellman and Schofield, 2007) available on the SEDAR website (www.sedar.com).

13.1

Redport Ltd 2005 aircore drilling

Redport completed aircore drilling over the LMUP on exploration licences E53/1099 and E53/947 during October and November 2005. Redport drilled a total of 590 (LMAC0001-0590) holes for 4,982.5 m. At the time this represented 37% of the resource database by hole or 40% of the data set by metres (Princep, 2006). The majority of aircore drilling – completed on a spacing of 400 mN by 100 mE (MGA94) – infilled historical exploration and evaluation drilling completed between 1972 and 1979. This brought the nominal drill spacing to 200 mN by 100 mE within the area covered by E53/1099. In some areas of poorer drill coverage, drillhole spacing was closed down to 100 mN x 100 mE. In addition, two single lines, one north-south and one east-west, of 50 m-spaced holes were completed on 310,000 mE and 6,991,980 mN, respectively (see Figure 13-3). The aircore drilling was completed by Wallis Drilling Pty Ltd of Perth using a Wallis Mantis 75 track-mounted (Muskeg) aircore top-drive drill rig with HQ diameter reverse circulation drill rods and conventional blade bit. The capacity of the compressor was rated at 150PSI:160CFM. All drilling was carried out on a 12-hour day shift basis. Metre rates generally averaged 130 m (16 holes) per day. All holes were drilled vertically. Hole depths varied between 6 and 12 m, averaging 9 m. Each hole was cased with Class 9, 80 mm diameter PVCu (PVC) at the time of drilling, ensuring they remained accessible for radiometric logging at a later date. In all but a few instances, drillholes were successfully cased to the end-of-hole depth. All PVC drill collars were capped with a standard PVC end cap. The drillhole number was written on the PVC casing and the underside of the PVC end cap. In addition, the drillhole number and end-of-hole depth were recorded on the wooden grid peg located at each hole location. Drill samples were collected nominally at every half-metre downhole. The entire sample was collected directly from beneath the cyclone into a green plastic bag. Samples were not split. Wet samples were allowed to settle before the excess water was decanted off. All drill samples were collected from each drill site and relocated to a central sample bag farm located on the western edge of the lake surface. Each hole was logged for colour, major lithology, minor lithology, water and scintillometer reading.

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All holes were routinely radiometrically logged by Redport personnel using a portable gamma logging system. Only a few holes could not be logged below mineralised zones.

Figure 13-1: Wallis Drilling’s Mantis 75 track-mounted drilling rig used to carry out resource drilling by Redport Ltd in 2005

13.2

Mega Uranium Ltd 2007/2008 drilling

13.2.1 Aircore drilling Aircore drilling was completed in two campaigns – December 2007 to February 2008 and June 2008 to August 2009. A total of 11,677 m was completed in 794 aircore holes. The overall objectives of the resource evaluation drilling programmes undertaken by Mega during 2007/08 were the following: To bring the majority of the +200 ppm U3O8 portion of the Hellman and Schofield (2007) estimated Inferred Resource to an Indicated level of confidence as the required level on which to undertake subsequent feasibility studies. To further investigate the geological controls on the uranium mineralisation in order to develop an adequate geological model to constrain the resource estimation process. To verify the quality of the historic and Redport 2005 drilling including the radiometric logging, assaying, geological logging and bulk density measurements.

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The first campaign (holes LMAC0606-1137) closed down the overall drill spacing (inclusive of historical and Redport 2005 drillholes) to 100 mN by 100 mE within exploration licences E53/947 and E53/1099, and prospecting licences P53/1259-1261. Drill coverage was restricted to the + 200 ppm U3O8 grade panels of the Hellman and Schofield (2007) Resource. The second drill campaign (holes LMAC1138-1353 and LMAC1370-1415) provided contemporary drill coverage to the southern portion of the resource within tenements P53/1259-1261 and the southern extent of the main airborne radiometric uranium channel response within E53/1210. This area of the Inferred Resource had not been previously drill tested by Redport during 2005 as they did not hold the licences at the time. Aircore drilling was carried out on a nominal 200 mN by 100 mE spacing, infilling historic 1970s drilling completed mainly by BP Minerals. All aircore drilling was undertaken by National Drilling of Kalgoorlie using a purpose-built rubber track-mounted (Morooka vehicle) KL-150 drill rig. National Drilling constructed the rig in Kalgoorlie in 2007 at Mega‟s request. The rig onboard compressor was rated at 150PSI:300CFM. All holes were drilled employing an HQ diameter reverse circulation rod string with a 54 mm ID inner tube and a conventional blade bit with a nominal 90 mm OD. All holes were drilled vertically to a nominal depth of 12 m. Downhole surveys were considered unnecessary for such short drillholes as deviation was assumed to be negligible. Each hole was cased with Class 9, 80 mm diameter PVC and capped at the time of drilling so that they could be radiometrically logged at a later date. All holes were successfully cased past the zone of mineralisation (usually 2 - 5 mbgl). In most instances drillholes were successfully cased to the end-of-hole depth. All PVC drill collars were cut down to approximately 0.5 m above the surface and capped with a standard PVC end cap. The drillhole number was written on the PVC casing and the underside of the PVC end cap. In addition, the drillhole number and end-of-hole depth was recorded on the wooden grid peg located at each hole location. All drilling was carried out on a 12-hour day shift basis. Daily metre rates generally averaged between 180 and 200 m (15 - 17 holes) per day. Drilling conditions were generally dry for the first rod (3 m) and thence generally wet to moist for the remaining 3 to 12 m. Significant amounts of water injection were used on many of the holes (up to 2,000 l/day or 200 l/hole), particularly during drilling of sticky clays, which were found to be difficult to lift. The first sample after a rod change was commonly more saturated with water due to ingress of water into the hole. Due to the relatively unconsolidated nature of the surficial gypsum and lake muds, the first interval (0 - 0.5 m) of each hole generally exhibited poorer sample recoveries. Owing to the thin and horizontal nature of the uranium mineralisation and sedimentary host units, it was decided to conduct all aircore drill sampling on a nominal half-metre interval basis. This ensured consistency of sample interval and a similar sample size with the previous aircore drilling undertaken by Redport in 2005 and also provided more accurate location of geological contacts, thereby improving the accuracy of subsequent modelling of geological domains used in resource estimation. Half-metre samples were collected directly from the cyclone into a green plastic bag without splitting. Saturated samples were allowed to settle before the excess water was decanted off.

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Qualitative information of sample recovery and water content was recorded by Mega field assistants during drilling. All samples were tagged with a plastic bar coded cattle tag for later identification („LM‟ sample prefix). This information was recorded at the drill site into sample record books. These data were then entered and validated on site prior to loading into the relational drillhole database. Table 13-1: Summary statistics for resource infill aircore and sonic drilling programmes completed by Mega during late 2007 and 2008 Programme

Hole numbers

Holes

Metres

Samples

Aircore (Dec 2007 - Feb 2008)

LMAC0606 - 1137 LMAC1138 - 1353 LMAC1370 - 1415 LMSC001 - 042

532

6,378

12,755

262

5,299

8,213

42

618

1,236

Total

836

12,295

22,204

Aircore (Jun - Aug 2008) Sonic Core (Jan - Mar 2008)

Figure 13-2: National Drilling’s Morooka track-mounted KL-150 aircore rig used to carry out resource drilling

All drill samples were collected from each drill site and relocated to the central bag farm previously established by Redport in 2005. The existing fencing was extended to enclose all aircore half-metre samples collected in 2007/08 to prevent access by wildlife. All holes were routinely radiometrically logged by Borehole Wireline Pty Ltd of Adelaide using a portable wireline gamma logging system. A few holes could not be logged below the mineralised section.

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U3O8

Figure 13-3: Drillhole location plan of Mega 2007/08 and Redport 2005 drilling with current resource limits

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13.2.2 Sonic core drilling A 42 hole sonic core drilling programme totalling 618 m was undertaken between January and March 2008. Sonic core holes „twinned‟ a total of 18 Redport 2005 aircore drillholes. The main objectives of the sonic core programme were the following: To verify the quality of both the Redport 2005 and the Mega 2007/08 aircore drilling (chemical assays and radiometric logging) To verify the validity of downhole radiometric logging as the main determinate of U3O8 grade To collect suitable material on which to conduct initial detailed metallurgical testwork To collect suitable material on which to conduct definitive disequilibrium testwork To collect suitable material on which to collect additional bulk density information To construct a number of suitable water bores on which to perform initial pump testing The drilling was undertaken by Boart Longyear (Sonic Division) of Perth using a track-mounted Minisonic drill rig. Drilling was undertaken at a total of 18 locations throughout the identified resource area (see Figure 13-3). At each location, both a PQ and SQ diameter sonic core hole twinned an existing Redport 2005 aircore hole. PQ holes were sited nominally at 5 m west and SQ holes at 2 m east of the existing Redport aircore hole. An additional SQ diameter hole was drilled at six of the locations and constructed as a water bore. PQ (87.0 mm diameter) core holes were drilled using the split barrel system, which uses a 1.5 m length polycarbonate liner tube. Core runs were nominally drilled to 1.5 m (length of the core barrel). The sonic drilling methodology involves drilling of a core run immediately followed by the advancement of steel drill casing. This ensures the hole remained open and could be cased to its terminal depth prior to retrieving the steel drill casing. SQ (129.7 mm diameter) core holes were drilled employing the conventional SQ sonic core barrel (3 m in length) with samples collected into flat laywrap sections and laid out in aluminium core trays. Core runs were nominally 3.0 m (length of the core barrel).

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Figure 13-4: Boart Longyear’s Minisonic track-mounted sonic core drill rig used by Mega in 2008

All holes were routinely radiometrically logged by Borehole Wireline of Adelaide Pty Ltd using a portable wireline gamma logging system.

13.2.3 Safety National Drilling and Boart Longyear complied with Mega‟s OHS and Radiation regulations. All drilling crew and Mega employees wore long trousers and long sleeved shirts. All personnel present at the drill rig were required to wear mandatory personal protective equipment (PPE) including hard hats, ear plugs, dust masks, safety glasses and safety boots. Radiation exposure was monitored during the drilling campaign. Each member of the drilling crew as well as Mega staff were equipped with personal dosimeters.

13.3

Drillhole collar surveying

All drillholes completed by Redport in 2005 and Mega in 2007/08 drilling programmes were surveyed by contract licensed surveyor MHR Surveyors of Perth using RTK-GPS. Mega engaged MHR in October 2007, March 2008 and October 2008. A summary of the drillhole collars picked up is provided in Table 13-2. Validation of the Redport drillhole collars identified a number of drillholes that had not been accurately picked up by RTK-GPS in 2006. These were relocated and accurately surveyed during October 2007. During the processing of these, Phil Richards of MHR discovered an error in the previously quoted RTK-GPS heights for the Redport 2005 drill collars surveyed in 2006. This error equated to a +13.42 mRL block shift in quoted AHD heights. A corrected file of all previously picked up Redport 2005 LMAC prefix holes was re-issued and used to update the drillhole database.

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Table 13-2: Drillhole collar survey pick-ups 2008 Pre-fix

Hole numbers

Date

LMAC

0206, 0473-0475

Oct 2007

LMAC

0602, 0606-1137

Mar 2008

LMSC

001-034, 036-042

Mar 2008

LMAC

1138-1353, 1370-1415

Oct 2008

All collar locations were referenced to Geocentric Datum of Australia 1994 in the x and y and Australian Height Datum (Ausgeoid adjusted). X and Y coordinates were quoted as Map Grid of Australia 1994 (MGA94) UTM Eastings and Northings. The location of the RTK base stations and tie-ins used during the surveying are shown below in Table 13-3. Accuracy of the RTK-GPS system is quoted to 0.05 m horizontal and 0.10 m vertical.

Table 13-3: Lake Maitland survey control points GDA94 (MGA Zone 51)

Point

AHD (Ausgeoid adjusted)

Description

Easting (m)

Northing (m)

Height (m)

MHR01

310193.186

6993275.750

472.976

Spike

MHR02

309684.866

6997410.711

473.413

Spike

SAM43

308414.452

6995899.261

473.900

SSM

13.4

Results

Drilling has verified the grade and thickness of previous Redport mineralised drill intercepts. An area of higher grade (>500 ppm eU3O8) uranium mineralisation than that predicted by the Hellman and Schofield Block Model (between 100 and 200 ppm eU3O8) has been identified in the south-eastern portion of the deposit. This area had sparse historical drill coverage having been previously drilled by Carpentaria at 400 mN x 100 mE. Additional mineralisation has also been defined outside the southern extents of the Inferred Resource. This north-east-trending zone of +500 ppm eU3O8 mineralisation is coincident with the major kopi headland located at the southern margin of the main Lake Maitland playa. Mineralisation in this area has been intersected at variable depths downhole but when adjusted for the effects of topography equates to the stratigraphic position of the main uranium deposit horizon between 2 and 5 m below the lake surface. Aeolian sand cover in this area has obscured the overall airborne radiometric surface response.

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Table 13-4: Summary of significant intercepts from Mega aircore drilling 2007/08 Drill programme Downhole - gamma intercept LMAC0606 - LMAC1137

LMAC1138-1353 LMAC1403-1415

Outer cut-off (ppm eU3O8)

200

100

Minimum width (m)

0.5

0.5

eU3O8 interval (m)

0.02

0.02

Minimum eU3O8 (ppm)

206

118

Maximum eU3O8 (ppm)

2112

600

Average intercept (ppm eU3O8)

594

231

0.50 m @ 5008 ppm eU3O8

1.28 m @ 600 ppm eU3O8

Average intercept width (m)

1.4

1.1

Minimum intercept width (m)

0.5

0.5

Maximum intercept width (m)

3.38

2.90

Minimum depth (m)

0.4

0.0

Maximum depth (m)

7.9

10.0

Best intercept

Note: Sample intervals and downhole widths approximate true thickness as mineralisation is essentially stratabound within horizontal layered host sediments.

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eU3O8

Hellmann and Schofield 2007 Resource Outline (100 ppm U308 cut-off) Hellmann and Schofield 2007 Resource Outline (200 ppm U308 cut-off)

Figure 13-5: Downhole gamma grade thickness image of Mega 2007/08 aircore drilling and Redport 2005 drilling Note: Aircore downhole gamma results compared with the Hellman and Schofield (2007) resource model 100 ppm and 200 ppm U3O8 cut-off block grades

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14

Sampling Method and Approach

14.1

Downhole gamma logging

14.1.1 Logging methods Downhole total gamma logging has been used as the routine method for the determination of U3O8 grade in the current resource estimate. Downhole logging of all aircore holes drilled by Redport in 2005 was completed by Redport personnel using company-owned wireline logging system and probes, under the direction of consulting geophysicist, David Wilson, of 3D Exploration. Logging of holes took place between October and November 2008 following completion of the drilling. Holes were logged using an AUSLOG Borehole Logging System built by Auslog Pty Ltd in 2005. The logging unit consists of a motorised winch with 450 m of single conductor cable (W450-1), an electronic unit (DLS5), and a total count radiometric probe (33 mm A075A gamma tool). The system is controlled by an Auslog, DOS based, software programme using a portable computer connected to the electronic control unit. Each hole was logged twice with raw counts acquired at nominal 1 cm intervals. The hole was initially logged down the hole at a speed of 10 metres per minute, and then logged up the hole at a logging speed of 2 metres per minute. Downhole gamma logging of all aircore and sonic drillholes completed in 2007 and 2008 by Mega was undertaken by Borehole Wireline Pty Ltd of Adelaide using a calibrated Geovista total gamma probe. Downhole logging was undertaken in two separate programmes, in February - March and October - November 2008, following the completion of the two Mega resource evaluation drilling campaigns. Logging runs were completed both downhole and uphole for each drillhole. Gamma logging data acquisition was only made for the uphole runs, which were undertaken at a speed of approximately 3 metres per minute. Raw counts were acquired at nominal 1 cm intervals and re-sampled to 2 cm upon conversion to standard LAS digital data. Water levels in drillholes were recorded in all instances at the time of downhole logging.

14.1.2 Calibrations and verifications Primary calibrations Both the Redport Auslog probes and Borehole Wireline Geovista probes were calibrated at the Department of Water, Land and Biodiversity Conservation calibration pits in Adelaide („Adelaide Models‟). Calibrations were carried out pre mobilisation of portable logging equipment to Lake Maitland and post demobilisation to Adelaide. Calibration involves logging test pits (AM1, AM2, AM3) with known grade and thickness to determine the response of the logging system and then calculating a Calibration („K‟) Factor, which gives the true grade (i.e. conversion of counts per second to equivalent U3O8). A total of two passes of the gamma probe is made in each model. From this, a system deadtime and a probe calibration („K‟) factor were calculated for the probe. The raw counts were converted to equivalent U3O8 (eU3O8) by application of the system deadtime, K factor and a hole size factor. The Redport Auslog probes were also calibrated for hole size correction in test pit AM7. Hole size corrections for each probe in water were measured in pit AM7. This pit has five different hole sizes through the same mineralised zone. The logged hole in each of the calibration pits AM1, AM2 and AM3 was water filled and 108 mm in diameter.

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Daily calibrations Gamma probe verifications using a gamma jig involved placing the jig apparatus on the gamma probe at a known and repeatable location on a daily basis. The gamma jig is constructed in such a manner as to allow the apparatus to be placed at the same position of the gamma probe and to keep the Cesium 137 source (5 micro-curies) at a fixed distance from the gamma scintillation detector. Daily differences between the background reading value (jig not attached) and the jig value were negligible throughout the duration of all logging programmes. Verification hole logging Both Redport and Mega regularly logged existing aircore holes that served as reference holes (measure of repeatability). These were logged numerous times during each logging programme to establish any daily variation and drift that may have occurred. Multiple repeat logging of reference holes during gamma logging programs demonstrated good repeatability. This suggests that the build-up of radon gas in drillholes is not significant. Between September 2005 and January 2006, repeat logs of LMAC0012 on different days show less than a 0.5% variation in the calculated grade x thickness. Repeat logging of LMAC0012 during Mega logging programmes is discussed in Section 16.2.1. Repeat logging Apart from re-logging of reference hole LMAC0012, Redport do not appear to have undertaken any repeat logging of holes during the 2005 logging programme. Mega completed repeat logging of a selected number of drillholes during both logging programmes to ensure that primary logging run measurements were repeatable. A number of Redport holes were also re-logged to provide a cross check against the measurements of the Auslog and Geovista probes. Repeat logging showed very repeatable logs compared with the primary logging run (see Section 16.2.2). Cased and uncased logging checks Both Mega and Redport completed logging runs inside the PVC casing string and in the open hole once the PVC casing had been removed to establish the attenuation effect of casing on the gamma response. Under most circumstances, the logging was performed on different days. During the February - March 2008 programme this check was performed for both aircore holes (Class-9 80 mm casing) and PQ and SQ sonic holes (Class-12 80 mm PVC casing). A total of 18 aircore holes and 14 sonic holes were logged cased and uncased to establish a casing correction factor. The October - November 2008 logging programme used the same model Geovista probe, and casing correction factors derived for the February - March 2008 programme were used to correct gamma measurements.

14.1.3 Calculation of eU3O8 All raw data were provided to consulting geophysicist, David Wilson, of 3D Exploration Perth. The final equivalent U3O8 (eU3O8) for each radiometric measurement is established as follows: Equivalent U3O8 values at 2 cm downhole interval are calculated as: Edited raw gamma counts per second (cps) x dead-time corrections x K factor x hole size correction x air/water correction x casing correction.

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The corrected data are then filtered using a three-point non-linear filter to remove single reading spikes followed by a triangular average filter of nine points to smooth the data. Data are then deconvolved to sharpen contacts and produce slightly more accurate grades. The deconvolution filter is effectively a high pass filter that improves the resolution of true grade and true thickness but also amplifies the statistical noise (Wilson, 2006). The linear and triangular filters are applied to reduce the amplification of statistical noise during the deconvolution. The deconvolution filter has been determined by comparing the measured and calculated radiometric response in the AM1 test pit using the method of Conaway and Killeen (1978). All data including raw counts, corrected counts, eU3O8 and filtered/deconvolved eU3O8 are provided at 2 cm downhole in a Microsoft Access database. The un-filtered/un-deconvolved eU3O8 was used for the purpose of grade estimation.

14.2

Geological logging

14.2.1 Redport Ltd Redport geologically logged each half-metre aircore drill sample for colour, lithology, water content and scintillometer reading. Upon taking control of the project, Mega observed that Redport had not undertaken geological reference chip tray sampling of the drilling completed in 2005. During a refurbishment and sample audit of Redport‟s bag farm, Mega collected chip tray samples of all available samples in 2007.

14.2.2 Mega Uranium Ltd All aircore and sonic drillholes were geologically logged. Upon relocation to the central bag farm each half-metre aircore drill sample interval was wet sieved and representative chips were collected in plastic soil trays. In the case of clay-rich intervals, each chip tray compartment was half filled with a rapidly washed sample and the remainder was filled with any lithic fragments that remained after thorough wet sieving of the sample. This was done to ensure that the composition of the interval was consistently classified according to standard rock units. The drill chips trays are currently stored at Lake Maitland. All half-metre intervals were geologically logged using standard project logging codes. Each half-metre interval was logged for hardness, HCl reaction, colour, oxidation, weathering, regolith, lithology, grain size, texture, structure, descriptors, mineral, mineral style and mineral abundance. The logging data were captured in a Microsoft Excel spreadsheet template which was later validated and imported into the relational drillhole database.

14.3

Sampling

14.3.1 Redport Ltd Downhole radiometric values were used to control samples selected by Redport for chemical assay. Sample intervals of the entire length of the drillhole were submitted for 23 of the 54 holes assayed. Only uranium mineralised intervals as determined by the downhole radiometric readings were submitted for chemical assaying in the remainder of the 31 holes. Review of the downhole chemical uranium by Mega shows that in a number of these „selectively‟ sampled holes, the first and/or the last sample interval returned above 100 ppm eU3O8. This should be considered when making geostatistical comparison of the radiometric and chemical data sets (see Section 16.1).

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14.3.2 Mega Uranium Ltd In order to make unbiased geostatistical comparisons with the downhole gamma logging data of the Mega drilling, samples of the entire length of drillholes (n=81 holes) selected by Mega for chemical analysis were submitted for analytical determinations. Comparison of equivalent U3O8 and chemical U3O8 is discussed in detail in Section 16.1.

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15

Sample Preparation, Analysis and Security

15.1

Sampling

15.1.1 Redport Ltd Sample details Drill samples were collected nominally at every half-metre downhole. The entire sample was collected directly from the rig cyclone into a green plastic bag. Samples were not split. Saturated samples were allowed to settle before the excess water was decanted off. No quantitative sample recovery data (i.e. sample weights) appear to have been collected by Redport. Hardcopy records of qualitative recovery information were located for 22 holes. Recovery was stated as a percentage. Quoted recoveries varied between 15 and 130%, averaging 70-80%. Recoveries were markedly poorer (10-30%) in the first half-metre interval (0 - 0.5 m) of the holes recorded. Visual inspection of Redport aircore samples by Hellman and Schofield were made in 2007 who concluded that in general, sample recovery was reasonable (Hellman and Schofield, 2007). Sample recovery data are not considered a critical issue because the resource estimates rely primarily on downhole gamma measurements. Chemical assaying Redport undertook a programme of chemical assaying of aircore drill samples to check the radiometric values of the 2005 aircore drillholes. A total of 611 (6%) entire half-metre aircore drill samples were submitted for analysis for U, Th, K, V, Sr and SO4 by ICP-MS and ICP-OES. Full details of sample preparation and analytical methods are provided in the previous Lake Maitland Technical Report (Hellman and Schofield, 2007). No significant analytical precision or accuracy issues or problems with laboratory sample preparation practices were identified.

15.1.2 Mega Uranium Ltd Drill samples were collected nominally at every half-metre downhole. The entire sample was collected directly from the rig cyclone into a green plastic bag. Samples were not split. Saturated samples were allowed to settle before the excess water was decanted off. Mega routinely recorded qualitative sample recovery. All samples submitted for chemical analysis were weighed using a set of accurate bench scales. Sample weights varied between 0.1 and 13.5 kg, averaging 4.0 kg. These weights compare with a theoretical half-metre 100% recovery weight of between 4.1 and 7.3 kg based on the bulk density determinations of the geologically logged lithologies (see Section 19.2.2). An assaying programme of the December 2007 - February 2008 resource evaluation aircore programme was undertaken to provide comparison with the downhole radiometric logging eU3O8 values. A total of 2,016 entire half-metre (i.e. unsplit) aircore samples (16%) from 83 holes were submitted for assaying in three separate batches to Actlabs Pacific of Perth. Mega submitted a total of six Redport 2005 aircore holes for comprehensive multi-element analysis (see Table 15-1) as part of a sighter metallurgical testwork programme (see Section 18.2) in late 2007. Mineralised intervals (~ ≥100 ppm) were submitted to Ultratrace by AMMTEC who carried out the testwork. Mega submitted the remaining intervals for these holes were submitted to Genalysis Laboratory of Perth for a similar suite of multi-element analyses (see Table 15-2).

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Dispatch of samples from site was carried out by Mega employees under the supervision of project geologists. Samples were freighted by road to Perth as „excepted packages‟ UN2910 under the statutory radiation transport requirements. Table 15-1 Metallurgical sighter test samples – Ultratrace multi-element analysis details Method

Elements and detection limits*

Cast with 12:22 flux to form glass bead – analysed by XRF Spectrometry

Nb2O5 (0.005%), P2O5 (0.001%), SrO (0.001%), V2O5 (0.001%), U3O8 (0.001%), ThO2 (0.001%)

4-Acid Digest/reflux (HF,NO3,HCl, HClO4) analysed by ICP Optical Emission Spectrometry (ICP- OES)

Cu (2), Zn (2), Co (5), Ni (5), Mn (2), P (50), Sc (1), V (5), Al (0.01%), Ca (0.01%), Na (0.01%), K (0.01%), S (50)

4-Acid Digest/reflux (HF,NO3,HCl, HClO4) analysed by ICP Mass Spectrometry (ICP-MS)

As (1), Ag (0.5), Ba (1), Be (0.1), Bi (0.1), Cd (0.5), Ga (0.2), Li (0.5), Mo(0.5), Pb(1), Sb (0.2), Sn(1), W (0.5), Ta (0.1), Y (0.1), Hf (0.2), Zr (1), Nb (0.5), La (0.1), Ce (0.1), Pr (0.02), Nd (0.05), Sm (0.05), Eu (0.05), Gd (0.2), Tb (0.02), Dy (0.05), Ho (0.02), Er (0.05), Tm (0.02), Yb (0.05), Lu (0.02), Th (0.1), U (0.1), Se (5), Rb (0.2), In (0.02), Te (0.2), Cs (0.1), Re (0.1), Tl (0.1)

Fused with Sodium Peroxide – melt dissolved in dilute HCl – analysed by ICP Optical Emission Spectrometry (ICP- OES)

B (20), Cr (50), Si (0.01%), Fe (0.01%), Mg (0.01%), Ti (0.01%)

Fused with Sodium Peroxide – melt dissolved in dilute HCl – analysed by ICP Mass Spectrometry (ICP-MS)

Ge (0.2), Sr (20)

Thermo-Gravimetric Analysers (LOI371, LOI1000)

LOI (0.00%)

Note: Detection limits quoted as ppm unless otherwise stated.

Table 15-2: Metallurgical sighter test samples – Genalysis multi-element analysis details Method

Elements and detection limits*

CALC

C-CO3 (0.01%)

HotAcidInd/IR

TOC+C (0.01%)

Ind/IR

C (0.01%)

A/MS

Ag (0.2), As (2), Ba (0.1), Bi (0.01), Cd (0.1), Ce (0.01), Co (0.1), Cs (0.05), Dy (0.01), Er (0.01), Eu (0.01), Ga (0.1), Gd (0.01), Hf (0.01), Ho (0.01), In (0.01), La (0.01), Li (0.1), Lu (0.005), Nb (0.05), Nd (0.01), Pb (2), Pr (0.005), Rb (0.05), Re (0.01), Sb (0.05), Se (2), Sm (0.01), Sn (0.1), Sr (0.05), Ta (0.01), Tb (0.005), Te (0.1), Th (0.01), Tl (0.02), Tm (0.01), U (0.01), W (0.1), Y (0.05), Yb (0.01), Zr (0.1)

A/OES

Al (50), Ca (50), Cr (5), Cu (1), Fe (0.01%), K (20), Mg (20), Mn (1), Na (20), Ni (1), P (50), S (50), Sc (1), Ti (5), V (2), Zn (1)

D/OES

B (50)

Note: Detection limits quoted as ppm unless otherwise stated.

Sample preparation comprised drying, oven drying to 110oC, crushing to nominal 25 mm using two-stage Jacques crusher, rotary split, pulverising of an approximate 1.5 kg split using a singlestage grind LM-2 (Cr-steel bowl). Pulps were analysed for a 56 multi-element suite as detailed in Table 15-3. Due to delayed turnaround times encountered, it was decided that two of the jobs (542 samples) that had already been sample prepped at Actlabs should be submitted to Genalysis Laboratory in Perth for analytical assaying. Pulps were submitted for multi-element analysis as detailed in Table 15-3 and Table 15-4.

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Table 15-3: Mega aircore - Actlabs analytical details Method

Elements and detection limits* Ag(0.05), Al(0.01%), As(0.1), Ba(1), Be(0.1), Bi(0.02), Ca(0.01%), Cd(0.1), Ce(0.1), Co(0.1), Cr(0.5), Cs(0.05), Cu(0.2), Dy(0.1), Er(0.1), Eu(0.1), Fe(0.01%), Ga(0.1), Gd(0.1), Ge(0.1), Hf(0.1), Hg(0.01), Ho(0.1), In(0.1), K(0.01%), La(0.5), Li(0.5), Mg(0.01%), Mn(0.5), Mo(0.1), Nb(0.1), Nd(0.1), Ni(0.5), Pb(0.5), Pr(0.1), Rb(0.1), Re(0.001), Sb(0.1), Sc(0.1), Se(0.1), Sm(0.1), Sn(1), Sr(0.5), Ta(0.1), Tb(0.1), Te(0.1), Th(0.1), Tl(0.05), Tm(0.1), U(0.1), V(1), W(0.1), Y(0.1), Yb(0.1), Zn(0.2), Zr(1)

HFICP/UT-4

Note: Detection Limits quoted as ppm unless otherwise stated.

Table 15-4: Mega aircore - Genalysis analytical details Method

Elements and detection limits*

AT/MS

Ag(0.2), As(2), Ba(0.1), Be(0.1), Bi(0.01), Cd(0.1), Ce(0.01), Co(0.1), Cs(0.05), Dy(0.01), Er(0.01), Eu(0.01), Ga(0.1), Gd(0.01), Hf(0.01), Ho(0.01), In(0.01), La(0.01), Li(0.1), Lu(0.005), Mo(0.1), Nb(0.05), Nd(0.01), Pb(2), Pr(0.005), Rb(0.05), Re(0.01), Sb(0.05), Se(2), Sm(0.01), Sn(0.1), Sr(0.05), Ta(0.01), Tb(0.005), Te(0.1), Th(0.01), Tl(0.02), Tm(0.01), U(0.01), W(0.1), Y(0.05), Yb(0.01), Zr(0.1)

AT/OES

Al(50), Ca(50), Cr(5), Cu(1), Fe(0.01%), K(20), Mg(20), Mn(1), Na(20), Ni(1), P(50), S(50), Sc(1), Ti(5), V(2), Zn(1)

Note: Detection Limits quoted as ppm unless otherwise stated.

External QA/QC Certified reference material (CRM) was included with all Mega laboratory submissions. The initial batch of samples included CANNMET CRM as was used by Redport. Subsequent batches of samples submitted to Actlabs and Genalysis included CRM sourced from OREAS in Victoria. Standards were included at the rate of 1 in every 20 samples. Table 15-5: Certified standards used by Mega Value U (ppm)

Value U (ppm)

Fusion ICP

Pressed powder pellet

OREAS 104

127

-

Crocker Well U Deposit, SA

OREAS 105

530

-

Crocker Well U Deposit, SA

OREAS 106

1142

1213

Crocker Well U Deposit, SA

OREAS 100a

135

130

Mt Gee IOCG-U Deposit (blended with barren rhyodacite material)

OREAS 101b

396

387

Mt Gee IOCG-U Deposit (blended with barren rhyodacite material)

OREAS CRM

Source material

Blank material sourced from the spoils of the exploration camp refuse pit was used to prepare coarse blank (3 kg) material to monitor carry-over during the sample preparation stage of laboratory analysis. Blanks were submitted at the rate of 1 in every 33 samples.

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Internal QA/QC Actlabs have routinely used OREAS CRM standards as laboratory analytical standards. Table 15-6: Certified reference material used by Actlabs Pacific OREAS CRM

Value U (ppm) Fusion ICP

Value U (ppm) 4 Acid ICP

Source material

OREAS 100a

135

130

Mt Gee IOCG-U deposit (blended with barren rhyodacite material)

OREAS 102b

662

638

Mt Gee IOCG-U deposit (blended with barren rhyodacite material)

OREAS 45P

2.4

2.4

Ferruginous soil, South Murchison Western Australia (blended with barren soil)

15.1.3 Adequacy of analytical procedures The analytical methods and laboratories used are considered appropriate. The sampling methods, chain of custody procedures and sample preparation procedures are all considered appropriate and are compatible with accepted industry standards. Genalysis Laboratory Services Pty Ltd is accredited under the National Association of Testing Authorities Australia (NATA), following demonstration of its technical competence, to operate in accordance with ISO/IEC 17025 (1999), which includes the management requirements of ISO 9002:1994. This facility is accredited in the field of Chemical Testing for the tests, calibrations and measurements shown in the Scope of Accreditation issued by NATA. Actlabs are currently in the process of achieving ISO 17025 accreditation through NATA.

15.1.4 Bulk density determinations The bulk density work on the calcrete material is described in detail in Section 19.

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Data Verification

The 2009 Resource estimation relies entirely on downhole total gamma radiometrics, as transformed to eU3O8. The current QA/QC analysis performed on chemical assaying is not comprehensive, but chemical assays are not the primary source of data for the resource estimate. They serve as a check and calibration tool to the radiometric data. Hellman and Schofield (2007) provides a detailed assessment of the QA/QC on the historical data and on the 2005 Redport data. This will not be repeated in this Report – only recent work is reported.

16.1

Review of downhole gamma vs chemical assays

At the request of Mega, SRK was asked to review a comparison of the gamma assay and chemical assay downhole data at the LMUP. It should be noted that some of this work has been repeated in the original QA/QC work on the data completed earlier in the programme. However, it does reinforce the view that, in general, the sampling is of a high quality. Based on an analysis of the assay data, it is SRK‟s opinion that whilst the slope of regression between two individual sample types is not high, there is little reason to doubt the overall quality and reliability of the downhole gamma eU3O8 data on which the resource estimate is based. The regression analysis in itself is not good for doing comparative analyses between two essentially different samples. In SRK‟s opinion, the moderate level of correlation seen between the two types of assay data on the scatter plots is to be expected considering: The relative physical size of samples represented Differences in sample type with differing sample supports That only the difference between individual samples, and not the resource or mineralised zones as a whole, is reflected SRK performed a number of different statistical and analytical tests in addition to correlation analysis to determine if the gamma analyses were representative of the mineralisation and grades recovered during drill sampling compared to the chemical samples. Based on these findings, SRK is of the opinion that the gamma sampling is representative of the grades likely to be encountered during mining at the level of classification currently attributed to the resource. Mega supplied SRK with a database (2,643 in total) of 0.5 m sample composites for the downhole gamma and equivalent chemical assays. Of the 2,643 samples, approximately 30% (690) came from the Redport drilling programme whilst the remaining 70% (1953) came from the Mega drilling programme. The data from the two drilling programmes were treated separately for the purpose of statistical analysis.

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Data were analysed using the following methods for comparison between sample types – chemical and gamma: Comparison of summary statistics (mean, standard deviation, etc.) of the different methods Scatter plot and regression analysis Histograms of grade distribution QQ plots to determine if the sample data sets were of the same population group Downhole comparative spatial analyses of the two sample types (chemical vs gamma) Prior to any statistical analyses, the actual uranium (U) values reported from the chemical ICP-MS data were converted to U3O8 values for direct comparison with the downhole gamma log eU3O8 values. A conversion factor of 1.179 was used.

16.1.1 Basic statistical analysis A comparative statistical analysis was made between the raw ICP-MS chemical uranium values, the converted U3O8 chemical (ICP-MS) values and the gamma log values for the 0.5 m composites for each drilling programme (Redport and Mega). The statistical results can be seen in Table 16-1 and Table 16-2. Table 16-1: Mega 0.5 m composites summary statistics Sample type

Arithmetic mean

Standard deviation

Coefficient of variation

Minimum value

Maximum value

Uranium ppm chemical

66.4

183

2.7

0

2882

U3O8 ppm chemical

78.3

216

2.7

0

3398

eU3O8 ppm gamma

70.9

207

2.9

0

4518

Table 16-2: Redport 0.5 m composites summary statistics Sample type

Arithmetic mean

Standard deviation

Coefficient of variation

Minimum value

Maximum value

Uranium ppm chemical

231

413

1.78

0

3554

U3O8 ppm chemical

272

487

1.79

0

4170

eU3O8 ppm gamma

229

407

1.77

0

2985

A standard set of frequency histograms showing the distribution of eU3O8 (gamma) and U3O8 (chemical) for each of the two data sets can be seen in Figure 16-1 and Figure 16-2. From a review of the statistical data alone, it is clear that the population distributions for the gamma vs chemical assays for both data sets are similar. Within the Mega data set, the mean values for the eU3O8 data (gamma) and the U3O8 data (chemical) are similar – the chemical data being slightly higher in the global mean grade, by approximately 10% (71 vs 78 ppm). The coefficients of variation (CVs) are also very similar (2.7 vs 2.9) – this clearly indicates that the samples are from a very similar population group. The CVs for both data sets are relatively high, indicating that there is high variability in the data – it is reasonable to expect this, given the inherent spatial variability. The chemical assays also show a slightly higher standard deviation as would be expected from a small volume sample, compared to the larger volume sample of the gamma assays (volume-variance effect). Within the Redport data set, the mean values of the eU3O8 data (gamma) and the U3O8 data (chemical) are slightly different – the chemical data being higher in the global mean grade by approximately 17% (272 vs 229 ppm). This may reflect more selective sampling practices adopted HERO/GLEE/GUIB/WILL/mool

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by Redport (see Section 14.3.1). The CVs are also very similar (1.79 vs 1.77). Again, this clearly indicates that the samples are from a very similar population group. The chemical assays also show a slightly higher standard deviation (487 vs 407) as would be expected from a small volume sample, compared to the larger volume sample of the gamma assays (volume-variance effect). In comparison with the Redport data, the Mega data set is clearly lower grade and less variable – with means of 70 vs 229 ppm and standard deviations of 207 vs 407 for the gamma data. Again, this is a result of selective sampling by Redport. With respect to the Mega data, the average grade and distribution are similar between sample types. At the level of the resource estimation (Indicated to Inferred) any individual difference in grade between two samples is to be expected, given the different sample types. This will have little impact on the global estimate or on the estimate of the resource blocks in the current estimation update as each block may have well over 50 individual samples in its estimate and therefore the average grade between the two sample types is likely to be similar. The average grade for the chemical samples reports slightly higher (by 10%) compared to the gamma samples indicating a possible low bias of the eU3O8 values. Assuming this bias is confirmed, the current estimation could prove conservative.

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U3O8

U3O8 eU3O8

eU3O8

Figure 16-1: Histograms comparing chemical and gamma composited data sets for the Mega drilling

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U3O8

U3O8 U3O8

U3O8

Figure 16-2: Histograms comparing chemical and gamma composited data sets for the Redport drilling

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16.1.2 Regression analysis A scatter plot and regression analysis of the two sample types and two data sets were also completed. It is extremely important to note that absolute similarity between any two samples is not necessarily indicative of the quality of one sample type compared to another. The two sample types – gamma vs chemical assay – differ in how the samples are taken and the sample sizes involved, as detailed in the following section: Gamma log samples These samples are collected geophysically by a calibrated downhole probe approximately every 2 cm which measures the decay products of uranium (Bi214 and Pb214). The amount of uranium is calculated by the ratio of daughter products present. The uranium values are quoted as eU3O8 –„e‟ refers to „equivalent‟. The gamma rays can penetrate around the annulus of the hole for a radius of some 0.35 m or more, giving a sample diameter of approximately 0.7 m. Once composited to 0.5 m, this gives the effective sample volume of some 0.2 m3 – approximately 300 kg of sample. Gamma log sampling can suffer the effect of disequilibrium (positive and negative). Previous analyses by SRK and other consultants involved in estimates have ruled out any significant disequilibrium effects, and SRK does not regard this to be an issue in the current study (see Section 19 for a summary of the work on disequilibrium). Chemical sampling By comparison, the chemical assays are taken from blade bit cuttings blown up the air core drill hole at 0.5 m intervals. Sample contamination or smearing of sample down the hole can occur in this process. This sample represents a volume of approximately 0.004 m3 at best – assuming 100% recovery. The entire sample was sent to the laboratory. Where sample mass was greater than 3 kg, the sample was split to give 3 kg for pulverisation. The pulverised sample was further sampled to generate the required 2 g analysis sample. Assuming no bias is present, this sample is then analysed in the mass spectrometer to provide the actual uranium (U) value (not U3O8). The actual U3O8 value is then derived using a factor (1.179). That individual samples may differ significantly in grade is not unexpected – considering the methodology and sample size difference (300 kg for gamma vs 4 kg chemical). It is highly unlikely that, even under the most controlled circumstances, any two samples taken over a specific interval by each method will yield the same value. A significant degree of variability is therefore to be expected. In general, a high value sample will be high value in both sampling methods, although the actual grade could vary significantly. Scatter plot and regression analyses for both the Report and Mega data were conducted to examine the correlation between individual sample types only. Direct comparisons were made for the U3O8 chemical and eU3O8 gamma 0.5 m composites. Figure 16-3 and Figure 16-4 show the scatter plot and regression analysis for the Mega and Redport data sets. The results of the analysis for the Mega data set showed only a moderate correlation between any two samples (gamma vs chemical) with a correlation coefficient R equal to 0.7 – a value of 1 indicates perfect correlation. While the values are reasonably well clustered at the lower grade, they become more scattered at higher grades. If a few high grade outliers were removed, this would improve the correlation markedly.

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The Redport data set again showed only a poor to moderate correlation between any two samples (gamma vs chemical) with a correlation coefficient R=0.6. It is clear from the plot in Figure 16-3 that the values are quite dispersed, compared to those in the Mega data. The presence of numerous high grade outliers has a large effect on the correlation statistics. In general, this is a much higher grade data set. As mentioned previously, whilst the scatter plots do not show good individual correlation of samples, this cannot be taken as direct evidence of poor sample quality by the gamma data in comparison to the chemical assays. The quantile-quantile plots discussed in Section 16.1.3 are probably more indicative of the relative quality of the two sample sets and their relationship to each other.

eU3O8

U3O8

eU3O8

eU3O8

U3O8

Figure 16-3: Scatter plot for Mega 0.5 m samples – downhole chemical vs gamma samples

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eU3O8

U3O8

eU3O8

eU3O8

U3O8

Figure 16-4: Scatter plot for Redport 0.5 m samples – downhole chemical vs gamma samples

16.1.3 QQ plot analysis A series of quantile-quantile (QQ) plots for the Redport and Mega data sets were undertaken to make comparative analyses of the gamma vs chemical sample data sets. A QQ plot is used to determine if two data sets come from populations that have common or similar distribution, and is good for comparing data sets of different sizes. Figure 16-5 and Figure 16-6 show the QQ plots for gamma and chemical data sets for Mega and Redport, respectively. In the case of the Mega data set, the data are fairly well aligned and the two data sets have similar distributions. In addition, the data plot close to the theoretical distribution, indicating that both the gamma and chemical assays come from a very similar population. In the Redport plot, the alignment is not as good, but still acceptable with the data points only slightly diverging. Again, this demonstrates that the gamma values and chemical assays are derived from a common population. In both the Mega and Redport data sets, the data only drift from the theoretical curve at higher grades well above the resource cut-off. From these analyses, it is possible to say that the gamma and chemical assays are closely related and come from the same population group with similar distributions. These QQ plots are important in being able to demonstrate that, whilst the individual samples may differ, the overall behaviour of the two data sets is distinctly similar – statistically speaking, the chemical assays show a higher mean overall, as previously discussed.

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It is worth noting that the resource estimate model is a series of blocks with each block representing a particular volume of a portion of the deposit. Mining decisions – for example ore and waste classification and schedule of feed grades – are based on the block grades and not on individual sample grades. The block grades are estimated using a statistical algorithm – in this case, Kriging – which relies on the local spatial variability of the data. In this case, some 40 to 50 data points may be used to estimate a single block. In this context, the use of QQ data to compare the data distributions is considered to be a more appropriate tool than the direct comparison of individual samples. eU3O8

U3O8

eU3O8

eU3O8

U3O8

Figure 16-5: QQ plot for gamma vs chemical assays – Mega data set

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eU3O8

U3O8

eU3O8

eU3O8

U3O8

Figure 16-6: QQ plot for gamma vs chemical assays – Redport data set

16.1.4 Downhole comparison Some four holes from each of the two data sets (Mega and Redport) were randomly chosen from various locations in the resource and not just a single area. For each hole, the gamma and chemical assays were plotted against downhole depth to determine the spatial correlation between the two sample types. This exercise has been carried out several times by SRK in the recent resource estimate QA/QC study and found to give similar results. Figure 16-7 and Figure 16-8 show the downhole plots for the selected Mega and Redport drillholes. The Mega holes selected (Figure 16-7) all have extremely good correlation between the chemical and gamma assays for the anomalous uranium mineralisation zones. Again, it should be noted that whilst some individual values vary, the comparison between the sample types is clearly very good in determining where the mineralisation is situated stratigraphically, and the general levels of tenor. The Redport holes also show a very good correlation – with the exception of LMAC0538. In this hole, the gamma log has a single peak slightly offset from the chemical assay – by 1 m. It is possible that the actual downhole logging depths may be slightly offset between the two sample types in this hole – otherwise, the fit is very good. It is clear from these data, previous such exercises (Gleeson, 2009; Hellman and Schofield, 2007) on different holes, and visual inspection in 3D of all holes, that the mineralised zones are co-incident between the two sample types, thereby reinforcing the reliability of the gamma logs in comparison to the chemical assays.

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eU3O8

U3O8

eU3O8

U3O8

U3O8 U3O8

eU3O8

U3O8 U3O8

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eU3O8

U3O8

U3O8

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Figure 16-7: Downhole graphs for Mega holes showing gamma assay values vs chemical assay values

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eU3O8

U3O8

eU3O8

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U3O8

eU3O8

U3O8 U3O8

eU3O8

U3O8

U3O8

U3O8 ppm

U3O8

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Figure 16-8: Downhole graphs for Redport holes showing gamma assay values vs chemical assay values

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16.1.5 Conclusions of comparative analysis In conclusion, the comparative study on the Redport and Mega data sets between the chemical and gamma samples showed: 1. The Mega and Redport data sets appear to be different only in terms of average grade and variability. This would be expected as the Redport holes come from a higher grade portion of the deposit compared to the Mega holes. 2. Whilst there is only a moderate level of correlation between the gamma and chemical assay values on an individual sample interval basis, this is not considered a major issue as: The QQ plots confirm the sample types to be from very similar populations. The downhole plots confirm that the two sample types outline the same anomalous or mineralised zones at similar grades. The statistics show that the two sample types have similar global means. Mining decisions are made at block level and not at the scale of an individual sample. 3. At the categories being used (Indicated – Inferred) it is unlikely that the difference in individual sample type grade would have a serious impact on global or individual resource block model grades due to the large numbers of samples being used to estimate the block or area. 4. The fact that the chemical samples appear to be slightly higher than the gamma grades means that the resource estimates could be considered slightly conservative in terms of grade. It is SRK‟s opinion that the moderate correlation of different sample types on an individual basis as outlined in the scatter and regression plots will not have any significant impact on the reliability or accuracy of the resource estimates carried out using the gamma assay data alone. However, it is good practice from a QA/QC perspective to continue the ongoing sampling using both gamma and check chemical assays throughout the evaluation programme.

16.2

QA/QC

Hellman and Schofield, as part of their 2007 resource estimate and review, undertook a review of the QA/QC procedures of the Redport sampling. Hellman and Schofield reported that the Redport downhole gamma logging and calibration procedures employed in the preparation of the radiometric data in the 2007 evaluation were of a high standard and adequate for the purposes for which these have been used. SRK carried out a review of the overall data quality based on the latest Mega and Redport drill sample data used in the resource estimate. Quality assessment of the data has been considered for: A single reference hole – LMAC0012 Repeat radiometric logging Twinned holes Chemical assays QA/QC results (blanks, standards, pulp duplicates)

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16.2.1 Assessment of drillhole LMAC0012 LMAC0012 was logged on at least seven occasions during the 2006-2008 data collection programme. To assess the data, the following data processing was carried out: The radiometric data were loaded into Datamine and the date the hole was logged was preserved. For each date, the radiometric data were: Checked to ensure there was an interval starting from 0 m. If there was no such interval, this was assigned, so that the FROM value is equal to the first downhole interval in the hole and the eU3O8 value is also the same as the first interval recorded. This was done so that the data compositing of each hole starts at 0 m. De-surveyed Composited to 0.5 m starting at 0 m – this is the same as the assaying interval Composited radiometric and assay data were merged into a single file containing BHID, FROM, TO, U3O8, eU3O3_241008 and so on. Table 16-3 shows a summary of the key U3O8 (or eU3O8) statistics for the hole. The table shows that all the radiometric data have a similar minimum, maximum and mean. By comparison, the assay data are similar, albeit with a lower maximum and a higher mean. Figure 16-9 plots the data as a line chart with the downhole interval on the X axis and the eU3O8 or U3O8 data on the Y axis. The chart shows a very good match of all the radiometric data, indicating that the repeatability of the probe over the 7 day period is very high. The assay data are very similar to the radiometric data – the most noticeable difference is a 0.5 m offset in the down limits of the mineralisation.

Table 16-3: Statistical summary of assays and each log for LMAC0012 Assay

241008

261008

281008

301008

311008

31108

41108

U3O8

eU3O8

eU3O8

eU3O8

eU3O8

eU3O8

eU3O8

eU3O8

Minimum

13

12

10

12

12

10

11

10

Maximum

1,170

1,485

1,484

1,474

1,475

1,474

1,454

1,480

Mean

262

232

232

231

230

231

231

232

Variance

168,521

197,937

195,174

193,329

192,803

194,460

189,420

195,273

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Lake Maitland reference hole LMAC0012 eU3O8 from 2008 radiometrics compared to assay data

1600

1400

1200

1000

241008 e U3O8

261008 e U3O8

281008 e U3O8

301008 e U3O8

311008 e U3O8

31108 eU3O8

41108 e U3O8

Assay U3O8

800

600

eU3O8 or U3O8 (ppm)

400

200

0 0

0.5

1

1.5

2

2.5

3

3.5

4

4.5

5

5.5

6

6.5

7

Downhole FROM

Figure 16-9: Line chart of radiometric and assay data for LMAC0012

16.2.2 Repeat logging To analyse the repeat data, the following data processing was carried out: The separate radiometric databases were combined into two databases: 1. An original database 2. A repeat database: In all cases, the hole number for repeats (ending with „_rpt‟) was corrected so that it matched the hole number in the collar data (for example LMAC1044_rpt was corrected to LMAC1044 in the repeat data set). Radiometric data were loaded into Datamine and de-surveyed. Data were checked to ensure there was an interval starting from 0 m. If there was no such interval, this was assigned so that the FROM value is equal to the first downhole interval in the hole and the eU3O8 value is also the same as the first interval recorded. This was done so that the data compositing of each hole starts at 0 m. Data were composited to 0.5 m starting at 0 m. This was done separately for the original and repeat data. The original and repeat assays were combined into a single file for analysis. Table 16-4 summarises the statistics for the 0.5 m original radiometric data and the 0.5 m repeat radiometric data. The table shows a very good repeatability. Table 16-4: Statistical summary of assays and repeats for all holes

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Original eU3O8

Repeat eU3O8

Minimum

3

2

Maximum

746

749

Mean

47

46

Variance

9,013

9,047

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Figure 16-10 shows a scatter plot of the original and repeat radiometric data. Generally the data lie on a y=x line indicating good repeatability with no bias. Appendix 1 shows scatter plots of the data separated by hole. Broadly the figures show: A group of holes with good repeatability for holes LMAC1044 – LMAC1053: These holes show modest scatter and slope of the regression close to 0.9. These holes used Geovita probes GR3348 for the original radiometric log and GR3827 for the repeat. A group of holes with excellent repeatability LMAC1162 to LMAC1231: These show very limited scatter and a slope of regression of 1. Both the original and the repeat used probe GR3827. Poor repeatability in hole LMAC1155 – using probe GR3827 for both the original and the repeat. As expected, the repeatability when measured with the same probe is better than using different probes. None the less, the general repeatability of the data is good and acceptable. Some further investigation into hole LMAC1155 is warranted.

Lake Maitland - Repeat Gamma Logs

800.00 700.00 eU3O8 R_eU3O8

Repeat eU3O8 eU3O8 (ppm)

600.00

Linear (R_eU3O8) (R_eU3O8)

500.00

y = 0.9873x + 0.4356

400.00 300.00 200.00 100.00 0.00 0.00

100.00 200.00 300.00 400.00 500.00 600.00 700.00 800.00 Original eU3O8 eU3O8 (ppm) Figure 16-10: Scatter plot of radiometric repeat data – all holes

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Twinned holes

Most of the Lake Maitland resource area is drilled at 100 m (east) by 100 m (northing). However, there are some twinned holes throughout the resource. To analyse the twin holes, data were simply composited to 0.5 m for each hole commencing at 0 m. This was then used to produce plots of downhole depth (X axis) vs eU3O8, dU3O8 (deconvoluted) and U3O8 (if available). Figure 16-11 shows an example of a plot; a complete set of plots are contained in Appendix 2. Overall the plots show a similar intersection width and grade for the holes. Hole

X

Y

Z

Year

LMAC1066

310,003

6,992,230

471

2008

LMAC0135

309,998

6,992,230

471

2005

Lake Maitland Twinned Holes LMAC0135 / LMAC 1066 4000

3500

3000

eU3O8 eU3O8

2500

LMAC1035

U3O8 U 3O8

LMAC1035

eU eU3O8 3O8

LMAC1035

dU3O8 dU 3O8

LMAC1066

eU 3O8 eU3O8

LMAC1066

dU 3O8 dU3O8

2000

1500

1000

500

0 0

2

4

6

8

10

12

14

Down HoleDepth Depth Downhole Figure 16-11: Twinned holes LMAC0135/LMAC1066

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QA/QC of chemical assays

SRK had access to the assay quality control report for the period 1/11/2008 to 1/04/2009, prepared by ioGlobal.

16.4.1 Blanks Blank assays received by Actlabs were practically all under control while Genalysis showed seven excessive grades out of 23 submissions. This is too high and suggests some contamination risks at the laboratory. It will need to be closely monitored in the future.

16.4.2 Standards In general, the accuracy of the standard assay measurements is good, well in statistical control. There is one issue with standard OREAS 45P, where totally different values are obtained. This suggests that there has been an error in the identification of the standard.

16.4.3 Pulp duplicates The results obtained for both Genalysis and Actlabs show good repeatability of the measurements with very few outliers. As mentioned previously, the Hellman and Schofield report of 2007 details the results of comparative analysis of standards, repeats and duplicates. The report summarises that no serious issues in relation to the sample data were raised and that overall the quality of the data is of a high standard. SRK‟s more limited review confirms this opinion.

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Adjacent Properties

The Lake Maitland deposit is located within a region of Western Australia with a currently reported endowment of 180 Mlb of U3O8 in calcrete style deposits. In addition to Lake Maitland, the North Eastern Goldfields province around the towns of Wiluna and Leinster host the Yeelirrie deposit (BHP Billiton), the Lake Way and Centipede deposits (Toro Energy) and the Hinkler Well deposit (U3O8 Limited). Exploration for uranium in Western Australia has a long history. The currently known uranium deposits in the North Eastern Goldfields were all originally discovered in the early 1970s following government initiatives to fly regional scale wide spaced magnetic and radiometric surveys. Subsequent work through the decade focused on defining early resources at these deposits. In 1984 the Australian Federal Government introduced the Three Mines Policy which effectively curtailed all uranium exploration in Australia. Following changes in government at both federal and state levels in 1996 and 2008, respectively, interest in uranium exploration in Western Australia has been blossoming with the first JORC compliant resource being published on many previously known deposits. Lake Maitland and its surrounding deposits display similar geological characteristics. There is over 400,000 km2 of arid to semi-arid Tertiary (since the Pliocene) drainage systems developed throughout Western Australia and in the northern Yilgarn. These systems cover extensive Archean granites and greenstones which can be weathered to depths of 300 m. The combination of suitable uranium source rocks (granites) and appropriate depositional environments have contributed to the concentration of calcrete uranium deposits in the area. The deposits are classified by their sedimentary deposition, either as valleys, playas or terraces, and all are in sequences of clays and sands mixed with chemical sediments. As described above, a number of recent resources have been published on properties close to Lake Maitland. These are: Lake Way and Centipede – owned by Toro Energy with a combined resource of 8Mt @ 618 ppm U3O8 for a contained U3O8 content of approximately 11 Mlb (measured and indicated) and 12Mt @ 502 ppm U3O8 for a contained U3O8 content of approximately 14 Mlb (inferred), both at a cut-off of 200 ppm. Toro have completed a pre-feasibility on the project and are proceeding to an Optimisation Study (source – Toro Energy ASX Release 12 June 2009). Dawson-Hinkler Well – owned by U3O8 Limited with an Inferred resource of 20.7 Mt @ 228 ppm U3O8 for a contained U3O8 content of approximately 10.4 Mlb (source – U3O8 Limited website). Yeelirrie – owned by BHP Billiton. Although historical resources are mentioned in regard to Yeelirrie, BHP Billiton have recognised that these do not meet current reporting standards and as such their current work programme is designed to establish a new resource. McKay and Miezitis (2000) provide a summary of the uranium industry in Australia and they reference historical (1982) Western Mining Company (WMC) estimates for Yeelirrie of 35 Mt @ 1,500 ppm U3O8 for a contained U3O8 content of approximately 115 Mlb. Yeelirrie is the most significant uranium deposit in the region. In addition to projects with resources, a number of companies are actively exploring on properties near Lake Maitland. Of particular note are Enterprise Metals and Venture Minerals, both of whom hold tenement positions which cover „up-stream‟ sections of the Lake Maitland palaeochannel. Their exploration to date has been limited aircore drilling returning moderate uranium grades. Further details can be found in press releases of the companies available on their websites (www.reveremining.com.au and www.ventureminerals.com.au, respectively).

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This section aims to illustrate the presence of a series of uranium deposits in the locale of Lake Maitland. All the projects have been the focus of recent development by their owners suggesting a new uranium mining centre will emerge in the Eastern Goldfields region of Western Australia.

Figure 17-1: Map of the region around Lake Maitland showing significant nearby uranium deposits

Table 17-1: Summary of adjacent resources Deposit

Operator

Ore (Mt)

U3O8 Grade (ppm)

Contained U3O8 (kt)

Contained U3O8 (Mlb)

Cut-off (ppm U3O8)

Yeelirrie

BHP Billiton

35.000

1,500

52.500

115.8

N/A

Lake Maitland

Mega Uranium Ltd

31.200

366

11.419

25.2

100

Lake WayCentipede

Toro Energy Ltd

20.200

550

11.072

24.4

200

Dawson-Hinkler Well

U3O8 Limited

20.600

230

4.690

10.3

150

AbercrombyMillipede1

Barrick, Norilsk/ MPI

3.410

660

2.258

5.0

200

110.410

–

81.939

180.7

–

Total

1

Historic resource published by Acclaim Uranium Limited

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18

Mineral Processing and Metallurgical Testing

18.1

Introduction

A metallurgical scoping testwork programme was completed in April 2009, targeting a base case process flowsheet for treatment of the Lake Maitland deposit. Based on a sighter leach test programme completed in December 2008, the overall programme was broken down into the following sections of test work: Characterisation based on individual profile and composite testing performed on samples obtained from a costean excavated adjacent to air-core hole LMAC0314 Comminution testing on a limited sample of competent calcrete Preliminary evaluation of beneficiation, based on individual and combined wet scrubbing and attritioning approaches Scoping stage alkaline leach testing to assess the impact of sulphate mineralisation on reagent consumption and identify the means to mitigate effects of sulphate on sodium carbonate consumption Scoping and optimisation stage flotation testing as a means to mitigate the impact of sulphate mineralisation and reduce reagent consumption Alkaline leach testing and optimisation based on combined flotation, ambient and elevated temperature leaching Refining process case studies examining resin in pulp (RIP) and counter current decantation/ion exchange (CCD/IX) approaches Assessment of slurry rheology and performance of solid/liquid separation pilot testing to mitigate the impact of clays on water balance and reagent consumption Targeted engineering testwork to generate specific design criteria pertaining to slurry rheology and settling characteristics Preliminary process flowsheet testing based on combined flotation and alkaline leaching Feasibility study test work will commence in August 2009 to incorporate: Unit process confirmation and optimisation Process variability testing to confirm applicability across the resource, spatially and by main ore type

18.2

Sighter testwork

A sighter testwork programme was completed in December 2008 to assess the applicability of alkaline carbonate leaching for treatment of the Lake Maitland resource.

18.2.1 Sighter testwork samples The sighter alkaline leach testwork programme was initiated in October 2007 and conducted at AMMTEC in Perth to provide initial sample characterisation and leach kinetic data. Sighter samples consisted of 27 x half-metre aircore samples derived from the Redport 2005 drilling. Sample selection was based on a review of the multi-element data of the 54 Redport aircore holes (n=611 samples) undertaken by Mega in 2007. HERO/GLEE/GUIB/WILL/mool

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Inspection of the multi-element data (K, Th, U, V, Sr, SO4, CO3), downhole radiometrics and geological logging identified different possible uranium mineralisation sub-types throughout the deposit. Based on the Redport Inferred Resource uranium grade cut-offs, a total of 16 different possible mineralisation sub-types were identified. The sample matrix used to identify potential uranium ore sub-types is shown in Table 18-1. The table shows the number of Redport 2005 drillholes sent for multi-element analysis (n=54 holes) that contain a particular sub-type. Table 18-1: Sample matrix used to select sighter testwork samples Lithology (From Redport 2005 logging)

U3O8 Grade (ppm) Low (100-200)

Medium (200-500)

High (>500)

Sulphate content

14

13

4

>1%

15

13

6

1%

10

20

6

1%

NIL

NIL

1

=100 ppm U3O8) were selected on the basis that they were representative of uranium grade, lithology, spatially (vertically and laterally) and chemically (principally Sr and SO4 content). As none of the material from the Redport holes previously submitted for assaying was available in a suitable form (bulk rejects and pulps), the closest neighbour drillhole was selected. Sample weights ranged between 1 and 6 kg. Further sample details are provided in Table 18-2 and Figure 18-1 below. Table 18-2: Sighter test sample locations Northing

Easting

Maximum depth (m)

Water table (m)

Number

U3O8 (ppm)

V2O5 (ppm)

LMAC0336

6995181

311501

7

0.65

4

502

362

LMAC0307

6994780

310501

7

1.12

4

915

570

LMAC0314

6994782

311200

7

0.85

5

398

265

LMAC0162

6993181

309098

9

3.35

3

541

335

LMAC0134

6992180

309999

9

1.20

8

1,195

519

LMAC0567

6992076

311099

9

0.25

3

146

215

27

709

399

Identification

Total

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Figure 18-1: Sighter test sample locations Note: Hellman and Schofield Resource Model 200 ppm U3O8 (Blue) and 500 ppm U3O8 (Red) outlines are shown.

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18.2.2 Chemical characterisation Head assaying and physical characterisation were performed on all samples and comprised: ICP assay (62 element) and LOI determination with comparative XRF analysis of: U3O8, V2O5, Nb2O5, SrO, ThO2 and P2O5. Total carbon, organic carbon and carbonate analysis on final solid residues. Mineralogical analysis via QEMSCAN to provide a quantitative modal data and assess metal and gangue deportment and liberation. Significant analytes are shown below; average, minimum and maximum assays across the sample suite indicate a wide variation in chemical composition. Total sulphur assays represent sulphate indicating high sodium carbonate consumption rates in an alkaline leach. Table 18-3: Significant analytes sighter sample ICP head analysis Element

Average (ppm)

Minimum (ppm)

Maximum (ppm)

Standard deviation

Si

136,479

50,400

248,000

56,897

Ca

91,359

5,100

175,000

51,396

Mg

86,224

29,900

117,000

21,286

Al

34,445

7,900

102,000

24,762

Na

19,690

4,500

37,400

7,944

Fe

18,590

4,400

54,300

12,734

S

10,541

1,550

55,500

13,772

Sr

9,236

180

51,600

15,434

K

7,524

2,200

17,900

4,140

Ti

1,590

300

4,800

1,100

U

627

52

2,010

581

Mn

279

62

866

179

V

222

80

475

120

Ba

178

36

372

89

P

117

Neg

400

107

Cr

112

Neg

250

58

B

51

Neg

120

47

Zr

33

12

96

39

18.2.3 Physical characterisation Physical characterisation comprised: moisture, wet and dry bulk density and specific gravity determinations. Moisture levels varied widely, and while some samples appeared to have dried over time, maximum moisture content up to 27% was detected in samples abstracted at depths of 5 m to 6 m. Water table levels shown above in Table 18-3, taken together with the hypersaline nature of local water, indicated that significant chloride should be expected in final leach solutions.

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Table 18-4: Physical characterisation by location Wet bulk density (t/m³) Identification

Dry bulk density (t/m³)

Solid

Loose

Compacted

Loose

Compacted

% Moisture

LMAC0336

1.187

1.437

1.220

1.413

5.0

Specific gravity 2.514

LMAC0307

1.196

1.405

1.177

1.371

6.6

2.584

LMAC0314

1.291

1.624

1.358

1.627

3.3

2.524

LMAC0162

1.170

1.363

1.012

1.295

12.6

2.733

LMAC0134

1.174

1.479

1.211

1.429

8.3

2.619

LMAC0567

1.178

1.425

1.217

1.372

3.9

2.489

Total

1.199

1.456

1.199

1.418

6.6

2.577

18.2.4 Mineralogy Quantitative analysis was performed via QEMSCAN, with the following main findings: Uranium and vanadium mineralisation was identified as exclusively carnotite. Sulphur was detected and identified as exclusively sulphate. Significant dolomite was detected in all samples with variable amounts of clay. Significant levels of strontium and sulphate were detected. Strontium was identified as celestine (strontium sulphate), while other sources of sulphate where represented by gypsum (calcium sulphate), visibly noted in „as received‟ samples, and other minor sources, possibly barite (barium sulphate). Selected samples contained significant clay fractions, predominantly as smectite and vermiculite with minor kaolinite. QEMSCAN samples were pulverised to 50% passing 50 µm and found to have a highly variable range of liberation at fine size. Carnotite locking character was described generally as binary particles predominantly associated with dolomite, vermiculite and quartz. Modal mineralogy is reported below showing average, minimum and maximum mass % over all intervals. Table 18-5: Modal mineralogy Mineral

Mass (%)

Carnotite

Average 0.2

Minimum 0.0

Maximum 1.2

Celestine

3.2

0.0

19.6

Gypsum & minor sulphates

1.0

0.2

3.3

Dolomite

38.7

0.2

10.0

Vermiculite

18.1

4.4

45.9

Smectite

18.6

1.4

61.3

Kaolinite

0.6

3.3

43.7

Quartz

14.5

0.4

86.8

Feldspars

2.2

0.0

9.3

Magnesiohornblende

2.1

0.2

6.9

Iron oxides

0.8

0.1

3.7

Total

100.0

-

-

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The distribution between dolomite and clays was assessed based on individual intervals and is presented below. A wide variation in mineralogy and physical properties, particularly rheology, is indicated, with some samples occurring as almost exclusively dolomite and some almost all clay. Total sulphate levels are also shown, also indicating maximal sulphate for samples comprising 20% to 40% dolomite.

100.00

50.00

90.00 80.00

45.00 40.00

70.00

35.00

60.00 50.00

30.00 25.00

40.00

20.00

30.00

15.00

20.00 10.00

10.00 5.00

0.00 0.00

20.00

40.00

60.00

80.00

y = -13.521Ln(x) + 78.762 R2 = 0.7636

% SO4 SO4

Mass % Clay

Clay, Dolomite & Sulphate Distribution 200ppm U3O8 Cut Off

SO4

SO4

0.00 100.00

Mass % Dolomite

Figure 18-2: Dolomite, clay and sulphate distribution

18.2.5 Uranium alkaline leach extraction A total of 25 kinetic leach tests were performed on 0.5 m sample intervals; samples derived from intervals LMAC0134 (2.0 – 2.5 m) and LMAC0567 (2.0 – 2.5 m) were excluded, given head grades below a 100 ppm U3O8 cut-off. Sample lots were sub-sampled and then prepared at a nominal P50 size of 50 µm, consistent with previous historical testing on the deposit prior to performing kinetic leach tests under the following conditions: Sodium carbonate leachate with bicarbonate as required maintaining a pH of 10 Leach temperature maintained at 90ºC (reflux) Kinetic samples taken at: 1, 2, 4, 8, 12, 24 and 32 hours retention No oxidant addition applied for initial sighter testing, Eh monitored during each test. Final uranium and vanadium extraction is presented in Figure 18-3 below for all tests, and the following conclusions were drawn: Uranium extraction averaged 86.8%, comprising 19 tests returning an average of 94.2% and six with significantly lower extraction, averaging 63.4%. Four of the six tests exhibiting low extraction used material from drillhole LMAC0314 and the remaining two tests used material from the top intervals of LMAC0336 and LMAC0307. Reference to Figure 18-2 shows that these holes are located in the northern area of the resource. Uranium extraction showed a slight increasing trend at higher head grade. The level of extraction achieved in the moderate oxidising conditions employed (+90 to +200 mV vs calomel) indicated the native oxidation state of uranium in carnotite as predominantly hexavalent and readily soluble in alkaline lixiviant.

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The impact of lower head grade, extent of liberation and native oxidation state exaggerates the variance in vanadium extraction compared to uranium.

Metal Recovery %

Final Metal Recovery 100.0 90.0 80.0 70.0 60.0 50.0 40.0 30.0 20.0 10.0 0.0

Uranium Vanadium

0

500

1000

1500

2000

2500

Head Grade (ppm)

Figure 18-3: Final metal extraction and head grade

Uranium and vanadium extraction are shown in Figure 18-4 below, plotted against the head assay ratio of U3O8 to V2O5 based on average ICP and XRF analysis. Correlation with vanadium extraction indicated carnotite with a range of composition. Visual examination of „as received‟ samples suggested different oxidation states for vanadium, based on yellow and green colouration due to vanadium oxidation state.

Metal Recovery (%)

Final Metal Recovery 100.0 90.0 80.0 70.0 60.0 50.0 40.0 30.0 20.0 10.0 0.0 0.00

Uranium Vanadium

0.50

1.00

1.50

2.00

2.50

3.00

3.50

4.00

Head Ratio Uranium to Vanadium (U3O8:V2O5) (U3O8:V2O5)

Figure 18-4: Final metal extraction and head grade ratio

A strong relationship was demonstrated between strontium head grade (sulphate) and final uranium extraction, as shown in Figure 18-5.

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Uranium Extraction and Strontium Head Grade 100.0

Uranium Recovery (%)

90.0

R2 = 0.9175

80.0 70.0 60.0 50.0 40.0 30.0 20.0 10.0 0.0 0

1

2

3

4

5

6

Strontium (%)

Figure 18-5: Uranium extraction and strontium head grade

The six samples from the northern drillholes, reporting significantly lower uranium extraction (averaging 63.4%) are evident in Figure 18-1. Based on QEMSCAN modal analysis, lower uranium extraction correlated directly with elevated sulphate concentration. The figures below represent respectively, the correlation between uranium leach extraction and total sulphur (sulphate) head assay (Figure 18-6) and the same correlation, split into contributions from sulphate in celestine and sulphate associated with gypsum and other minor sulphates (Figure 18-7). Impact of Sulphates on Uranium Extraction 60000

Total S (ppm)

50000 40000 30000 20000 10000 0 0.0

10.0

20.0

30.0

40.0

50.0

60.0

70.0

80.0

90.0

100.0

Uranium Extraction (%)

Figure 18-6: Uranium extraction and impact of sulphate

The data demonstrate that the low uranium extraction exhibited for samples in the northern area of the resource is attributable to the presence of high celestine concentrations (typically above 0.6%) and high gypsum concentrations.

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Impact of Celestine and Gypsum on Uranium Extraction 50000

45000

45000 S (ppm) Celestine

S (ppm) Celestine

40000

40000

S (ppm) Gypsum and Minor Species

35000

35000

30000

30000

25000

25000

20000

20000

15000

15000

10000

10000

5000

5000

0 0.0

10.0

20.0

30.0

40.0

50.0

60.0

70.0

80.0

90.0

S (ppm) Gypsum + Minor Species

50000

0 100.0

Uranium Extraction (%)

Figure 18-7: Uranium extraction and impact of celestine and gypsum

The trend shown in Figure 18-7 shows the relative split between sulphate contributions due to celestine and gypsum + minor sulphates and uranium extraction. The following comments are relevant: The group of 19 leach tests exhibiting acceptable extraction (averaging 94.2%) was taken mainly from drillholes at the southern area of the resource and contained lower sulphate. Total sulphate contained in this group was a mixture of celestine with higher concentrations of gypsum, possibly located and naturally concentrated at the surface. Sulphates consume sodium carbonate lixiviant directly resulting in significantly elevated reagent consumption. Gypsum (CaSO4) comprises 70.6% sulphate, while celestine (SrSO4) contains 52.3% sulphate. Within the group of 19 leach tests exhibiting acceptable extraction (and assuming complete dissolution of calcium sulphates), the halo of higher sulphate due to gypsum (and other minor sulphates) indicates a larger impact on sodium carbonate consumption due to the presence of gypsum as opposed to celestine. The group of six samples exhibiting significantly lower extraction (averaging 63.4%) and taken from the northern drillholes, show very high sulphate levels relative to the majority of samples. This sample group also shows a variable sulphate contribution due to elevated celestine and gypsum. Leach kinetic data have been presented below, grouped by drillhole and interval on the basis of final uranium extraction. The data were modelled to allow estimation of the plug flow retention time required to achieve 90% uranium extraction in each case.

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Table 18-6: LMAC0134 leach kinetic data Head grade (ppm)

Uranium kinetic data

Interval (m) U3O8

V2O5

Sr

Total S

90% extraction (hours)

32 hour extraction (%)

5.5 – 6.0

174

463

8,353

5,400

> 32

89.6

2.5 – 3.0

359

188

240

2,600

8.3

91.0

3.0 – 3.5

580

259

295

3,500

4.0

95.4

5.0 – 5.5

1583

874

260

4,850

11.3

96.1

3.5 – 4.0

1263

524

285

5,050

8.3

97.7

4.0 – 4.5

1882

732

308

4,850

9.2

97.8

4.5 – 5.0

1972

779

308

5,100

7.9

97.8

Uranium Recovery (%)

LMAC0134 Uranium Leach Kinetic Data

100.0 90.0 80.0 70.0 60.0

134 5.5-6.0 m 134 2.5-3.0 m 134 3.0-3.5 m

50.0 40.0 30.0 20.0 10.0 0.0

134 5.0-5.5 m 134 3.5-4.0 m 134 4.0-4.5 m 134 4.5-5.0 m

0

5

10

15

20

25

30

35

Retention Time (hr)

Figure 18-8: Alkaline leach kinetic data LMAC0134

LMAC0134 Leach extraction was generally acceptable, achieving an average final extraction of 95.1%. Retention time to achieve a target of 90% extraction averaged 8.2 hours, apart from the 5.5 m – 6.0 m interval, where elevated celestine levels are evidenced by strontium assay. Interval 5.5 m – 6.0 m tested with a lower final extraction of 89.6%, as a result of elevated celestine and lower head grade. Total sulphur assay taken together with strontium assay indicate possibly lower gypsum content than other samples tested.

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Table 18-7: LMAC0307 and LMAC0162 leach kinetic data Location Hole

LMAC0307

LMAC0162

Head grade (ppm)

Uranium kinetic data

Interval (m)

U3O8

V2O5

Sr

Total S

90% extraction (hours)

32 hour extraction (%)

2.0 – 2.5

246

502

16,631

12,500

> 32

78.7

2.5 – 3.0

518

387

1,269

3,950

24.0

94.4

3.5 – 4.0

557

534

426

3,600

2.9

98.1

3.0 – 3.5

1,649

747

1,269

7,000

6.1

98.3

4.5 – 5.0

100

155

263

2,300

> 32

88.6

5.0 – 5.5

558

306

260

1,550

-

93.8

5.5 – 6.0

710

501

230

2,400

9.3

96.4

LMAC0307 Uranium Leach Kinetic Data 100.0

Uranium Recovery (%)

90.0 80.0 70.0

307 2.0-2.5 m

60.0

307 2.5-3.0 m

50.0

307 3.5-4.0 m

40.0

307 3.0-3.5 m

30.0 20.0 10.0 0.0 0

5

10

15

20

25

30

35

Retention Time (hr)

Figure 18-9: Alkaline leach kinetic data LMAC0307

LMAC0307 Leach extraction was generally acceptable, averaging 96.9% for all intervals apart from interval 2.0 m – 2.5 m, where elevated sulphate as a result of elevated celestine and lower uranium head grade resulted in a final extraction of 78.7%. An overall trend toward high extraction at depth is evident, tracking lower strontium and higher uranium head grade. LMAC0162 Trends for LMAC0162 were generally similar to that shown for LMAC0307.

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Table 18-8: LMAC0314 leach kinetic data Head grade (ppm)

Uranium kinetic data

Interval (m)

U3O8

V2O5

Sr

Total S

90% extraction (hours)

32 hour extraction (%)

1.5 – 2.0

152

189

38,176

52,900

> 32

47.2

2.5 – 3.0

439

293

51,548

24,800

> 32

55.9

2.0 – 2.5

377

293

49,953

25,600

> 32

57.8

3.0 – 3.5

570

336

34,265

23,400

> 32

70.0

3.5 – 4.0

205

170

6,995

7,750

37.5

88.5

LMAC0314 Uranium Leach Kinetic Data 100.0 Uranium Recovery (%)

90.0 80.0 70.0

314 1.5-2.0 m

60.0

314 2.5-3.0 m

50.0

314 2.0-2.5 m

40.0

314 3.0-3.5 m

30.0

314 3.5-4.0 m

20.0 10.0 0.0 0

5

10

15

20

25

30

35

Retention Time (hr)

Figure 18-10: Alkaline leach kinetic data LMAC0314

LMAC0314 Leach extraction averaged 63.8%, while initial leach rates were significantly lower than for other samples tested. Elevated strontium assays indicate the highest levels of celestine were derived from this area and four out of the six leach tests returning low extraction originated from LMAC0314. Leach extraction tended to improve with depth, which is consistent with results found in other test sites.

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Table 18-9: LMAC0336 and LMAC0567 leach kinetic data Location

LMAC0567

Uranium kinetic data

Interval (m)

U3O8

V2O5

Sr

Total S

90% extraction (hours)

32 hour extraction (%)

1.0 – 1.5

256

180

38,176

55,500

> 32

71.1

1.5 – 2.0

645

397

51,548

8,300

7.7

92.3

2.0 – 2.5

314

440

49,953

3,550

4.2

93.5

2.5 – 3.0

486

401

34,265

4,300

6.8

95.2

2.5 – 3.0

156

194

244

3,600

10.8

92.6

3.0 – 3.5

161

298

191

8,600

6.6

92.3

Hole

LMAC0336

Head grade (ppm)

LMAC0336 & LMAC0567 Uranium Leach Kinetic Data 100.0

Uranium Recovery (%)

90.0 80.0

336 1.0-1.5 m

70.0

336 1.5-2.0 m

60.0

336 2.0-2.5 m

50.0

336 2.5-3.0 m

40.0

567 2.5-3.0 m

30.0

567 3.0-3.5 m

20.0 10.0 0.0 0

5

10

15

20

25

30

35

Retention Time (hr)

Figure 18-11: Alkaline leach kinetic data LMAC0336 and LMAC0567

LMAC0336 Leach extraction averaged 88.0% as a result of lower extraction at 71.1% in the top interval tested. An average of 93.6% extraction was achieved for lower intervals down to 3 m depth. Strontium assays are elevated throughout, while higher sulphate assays occurred in the top interval where extraction reported lower. The magnitude of strontium and total sulphur assay appears inconsistent compared to other samples, although basic trends remained consistent showing increased extraction with depth and lower average total sulphate. Some instances of intermixing (i.e. smearing) in sampling between intervals has been suggested by the client. It is not known whether this was the case for LMAC0336. LMAC0567 Trends for LMAC0567 were generally similar to that shown for LMAC0336.

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18.2.6 Alkaline leach carbonate consumption Sodium carbonate consumption rates were noted as extremely high and variable, dependent on total sulphate concentration. Figure 18-12 summarises sodium carbonate consumed against total sulphur (sulphate) assay. Impact of Total Sulphate and Sodium Carbonate Consumption Rate y = 0.0031x + 10.489 R2 = 0.9431

Sodium Carbonate Consumption (Kg/t)

200.0 180.0 160.0 140.0 120.0 100.0 80.0 60.0 40.0 20.0 0.0 0

10000

20000

30000

40000

50000

60000

Total Sulphur (ppm)

Figure 18-12: Alkaline leach sodium carbonate consumption rate

Reference to sample head assay data presented in Table 18-10 shows average, minimum and maximum total sulphur (sulphate) assays over the sample suite. The correlation shown in Figure 18-12 indicates alkaline leach sodium carbonate consumption rates as shown below. Table 18-10: Estimated leach sodium carbonate consumption rates Unit

Average

Minimum

Maximum

Standard deviation

Total sulphur assay

ppm

10,541

1,550

55,550

13,772

Sodium carbonate consumption

kg/t

43.2

15.3

182.5

53.2

18.2.7 Chloride content Reference to Table 18-11 shows that samples tested were typically derived below the water table and, in view of the hypersaline nature of the water in this area, it was expected that leach solutions should contain significant levels of soluble chlorides. Table 18-11 shows chloride levels from north to south of the resource. Levels shown indicate that CCD chloride washing will be required ahead of final metal extraction, purification and precipitation stages.

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Table 18-11: Soluble chloride in leach solutions Hole

Interval (m)

Average chloride (mg/l)

Average chloride (g/l)

Northing

Easting

LMAC0336

1.0 – 3.0

28,165

42.24

6,695,181

311,501

LMAC0307

2.0 – 4.0

17,778

26.67

6,994,780

310,501

LMAC0314

1.5 – 4.0

17,994

26.91

6,994,782

311,200

LMAC0162

4.5 – 6.0

5,341

8.01

6,993,181

309,098

LMAC0134

2.5 – 6.0

13,117

19.68

6,992,180

309,999

LMAC0567

2.5 – 3.5

14,322

21.48

6,992,076

311,099

18.2.8 Conclusions Based on sighter leach testing, the following conclusions were made and formed the basis for scoping level testwork reported below in Section 18.3. Sample characterisation Chemical and physical characterisation indicates a wide range of chemical variability with typically high levels of sulphate, implying elevated sodium carbonate consumption rates in an alkaline leach. Samples also contained significant silica fractions. Silica can act as a source for sodium carbonate consumption. Where carbonate consumption due to silica is an issue, sodium carbonate can be replaced by the use of ammonium carbonate to reduce consumption due to silica. The deportment of clay within the resource indicates that a substantial proportion of silica should reside in the ultra-fine fraction (typically below 12 µm). High soluble chloride levels in the ore also show that a chloride wash (CCD) stage will be required ahead of final solution purification and product precipitation stages. While site water could be considered for the leach stage, good quality water will be required for solution stage processing. Installation of a reverse osmosis plant (RO plant) may be required in the event that suitable quality and quantity of water cannot be harvested or reclaimed from bores. Mineralogy Alkaline leaching is typically preferred for ores containing in excess of 7%–10% carbonate (as calcite). Lake Maitland ore (as dolomite and clay) contains typically 30%–55% carbonate, indicating that alkaline leaching is economically preferred over acid leaching with sulphuric acid. Sulphate was identified as celestine (strontium sulphate), typically occurring as a deposit on the surface of calcrete and also within the clay fraction. Gypsum (calcium sulphate) was also visibly evident in samples tested and occurred as whole crystals nominally 20 mm wide and of variable length. Taken together, gypsum and celestine act as sources of sodium carbonate consumption and will need to be removed ahead of an alkaline leach stage to reduce sodium carbonate consumption. Significant clay fractions were identified via QEMSCAN modal analysis. The levels of clay identified were highly variable and indicate a wide range in materials handling and slurry rheology properties. Given that the ore is predominantly clay and dolomite, it is concluded that treatment may require the separate consideration and treatment of dolomite and clays. The levels of clay measured will likely impact on the ability to achieve high underflow densities during CCD washing and slurry thickening stages. Impact here will also add to sodium carbonate loss, given that CCD washing and solution recovery will be required

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irrespective of the use of Resin in Pulp or Ion Exchange. The use of centrifuging or a combination of dewatering cyclones and/or slimes centrifuging and CCD washing may need to be considered to improve final tailings wash efficiency and sodium carbonate recovery. Clay levels and ability to achieve high thickener underflow density will impact on the ability to operate an Ion Exchange recovery circuit and on this basis Resin in Pulp may be preferred economically. Uranium leach extraction Sighter leach testing was undertaken without addition of oxidant and the number of tests returning acceptable uranium extraction within a nominal 32-hour leach indicates uranium predominantly in the hexavalent state and readily soluble in alkaline carbonate media. Leach kinetics, particularly initial leach rates, were reduced for samples with high levels of sulphate; celestine and gypsum form the two predominant sources of sulphate in samples tested to date. Impact of leach kinetics is seen to be as a result of two factors, namely the possibility of a lower oxidation state for some samples returning lower extraction and as a result of high sulphate content. On the basis that celestine occurs as a surface deposit in close association with carnotite, it is inferred that lower leach rate (particularly initial leach rate) is a result of either a local reduction in sodium carbonate concentration due to the close association of celestine and carnotite and/or occlusion of carnotite by insoluble strontium carbonate, precipitated during reaction and consumption of sodium carbonate. Given this, it is concluded that the addition of oxidant and the liberation and removal of sulphates (particularly celestine) would improve overall leach kinetics. Oxidant addition could be considered based on gas sparging, although, given the level of clays, it is suggested that lower mass transfer rates may result, reducing the effectiveness of gas phase oxidant addition. The impact of clays on slurry rheology and water balance may require that oxidant addition be considered via dry powder addition and/or gas sparging. While some intermixing of intervals may have occurred during excavation, the general trend to increasing uranium extraction at depth may indicate that celestine and gypsum deport at the surface. On the basis that sulphate mineralisation has a negative impact on leach rate and final extraction, it is concluded that resource domaining will need to allow for strontium and total sulphur assay to assist in defining overall process operating cost and/or mining sequence. Sodium carbonate consumption Sodium carbonate consumption rates were estimated based on trends exhibited in sighter leach tests performed at a P50 size of 50 µm. An average consumption rate of 43.2 kg/t over the sample suite was estimated for the leach stage. This excludes any allowance for any additional loss due to the possible impact of clays on sodium carbonate washing and recovery, post leach. Reference to Table 18-10 demonstrates that the sodium carbonate consumption rate is expected to be high and extremely variable. It is reasonable to expect that sub-economic consumption would result if no action was taken to mitigate reagent consumption by removal of sulphate ahead of an alkaline leach stage. High sodium carbonate consumption occurs as a result of elevated sulphate levels due to the presence of both celestine and gypsum. Beneficiation to remove sulphates ahead of alkaline leaching would act to reduce reagent consumption rates and improve overall process economics.

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Celestine occurs as a surface deposit in close association with carnotite (possibly also impacting leach rate as a result), while gypsum occurs as large crystals. Gypsum was visibly evident in samples tested and known to be concentrated at the surface. Scalping of gypsum at the mine, via separate scrubbing to repulp clays and screening may act to reduce overall sodium carbonate consumption.

18.2.9 Recommendations The following recommendations are made with regard to further scoping testwork: Resource domaining activities should consider strontium and total sulphate assay to allow representation of processing parameters within the resource block model. On the basis that carnotite occurs as a surface deposit, the beneficiation of dolomite by scrubbing and/or attritioning may result in an upgrade and reduction in front end mass. Heap leach testing was not conducted during the sighter stage, given limited sample availability. The proportions of clay fraction suggest heap leach could only be considered on competent calcrete fraction and preliminary bottle roll testing is recommended to assess uranium extraction and sodium carbonate consumption compared to an agitated leach approach. The majority of samples returned acceptable leach extraction without oxidant addition. Given the impact of sulphates on leach extraction and possibility of variable uranium oxidation state, it is recommended that oxidant addition be applied in further leach optimisation. Elevated sulphate levels act to increase sodium carbonate consumption and may also reduce leach rate. Removal of sulphates via a reverse sulphate flotation should be considered, together with exploiting any ability to scalp coarse gypsum at the mine. Coarse competent dolomite fractions will require comminution ahead of a sulphate flotation stage. While scrubbing and/or attritioning is recommended, the fine nature of clays and likely negative impact of grinding this fraction ahead of a flotation stage may require that clay fraction bypass the comminution stage if milling preparation of dolomite is considered. Given that the ore also contains significant silica fractions, it is recommended that any impact of silica on sodium carbonate consumption rate be addressed once action is taken to mitigate the impact of sulphate. In the case where silica is shown to contribute to elevated sodium carbonate consumption, its replacement with ammonium carbonate could be considered, albeit at increased reagent cost. Highly variable proportions of clay indicate difficult materials handling and slurry rheology. Scrubbing should be considered as a means of effectively re-pulping ROM material and separating clay and dolomite fractions ahead of comminution and wet processing. Centrifugal pumping is applicable typically up to an 80 Pa shear limit and it is recommended that slurry rheograms be generated (forward and reverse scan) to assess pumping characteristics for fresh feed slurry as well as leach residue. Dilution and/or positive displacement pumping may need to be considered as an alternative to centrifugal pumping where slurry rheology is limiting. A measure of the maximum underflow density achievable is possible by noting the % solids at which an 80 Pa shear limit is achieved in slurry rheology testing. In the event this is shown to be low, e.g. 30% to 35% solids, it is recommended that alternative methods to CCD be considered. The parallel use of dewatering cyclones and/or centrifuging in conjunction with CCD washing is suggested.

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The size by size metal deportment needs to be determined in further testing to assess any possibility to de-slime the ore at an early stage, rejecting slimes and potentially improving overall slurry rheology. Since it is expected that the greater proportion of silica should occur in the fines, it is also considered that early de-sliming may act to mitigate any potential impact of silica on sodium carbonate consumption rate.

18.3

Scoping testwork

18.3.1 Introduction A metallurgical scoping testwork programme was completed in April 2009 targeting a base case process flowsheet suitable for treatment of the Lake Maitland uranium deposit. Scoping testwork was planned based initially on recommendations from sighter stage characterisation and leach testing (Section 18.2.9).

18.3.2 Scoping testwork samples A bench scale scoping testwork was conducted based on a 2 tonne bulk sample excavated from a costean adjacent air-core hole LMAC0314 (311200 mE, 6994782 mN). Sample selection was justified considering previous lower extractions and high carbonate consumptions reported from samples tested in this area (refer Table 18-8). The sample obtained also mimicked the expected average resource representation of clay to calcrete and typically contained a higher average sulphate grade than the expected resource average. The following testwork was performed: Preliminary comminution property characterisation Beneficiation testing to screen a number of potential routes for upgrading the feed Reverse flotation sighter and scoping testwork to mitigate impacts of sulphates Heap leach bottle roll testing for comparison with reverse flotation and agitated leaching Alkaline leaching to assess ambient pre-leaching of sulphates in conjunction with elevated temperature leaching Preliminary flowsheet testing incorporating reverse flotation and alkaline leach testing

18.3.3 Comminution testing Comminution testwork was conducted on competent calcrete fraction screened above 300 µm. Coarse competent calcrete was targeted in this case given that flotation sighter testing (reported below) indicated optimum flotation performance was achieved by treating the natural -300 µm fraction, obviating the need to mill combined clay and calcrete fraction. Six samples of competent calcrete were utilised for the following testing: Bond Crushing Index Bond Ball Mill Index Bond Abrasion Index The results provided in Table 18-12 below indicate a relatively low power consumption required to achieve appropriate size reduction. The abrasion index of the sample is classified as very low and indicates that the use of toothed crushers or mineral sizers should be possible.

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Table 18-12: Preliminary comminution indices competent calcrete Comminution test

Unit

Average

Maximum

Crushing Index

Kwh/t

6.2

9.5

Ball Mill Index

Kwh/t

8.6

9.2

Abrasion Index

–

0.02

–

Extensive comminution variability testing is yet to be conducted across the resource; however, a study of the resource lithology types showed that while there are 13 lithology types, the coarse, comprising the +0.5 mm particle size fraction, typically represents calcrete and mudstone with all others making up less than 5% of lithology types requiring comminution. On this basis, approximately 40% of the currently defined resource (200 ppm eU3O8 cut-off) naturally occurs in the fines (–0.5 mm) that are envisaged to bypass crushing and grinding unit operations.

18.3.4 Beneficiation testing Beneficiation testwork was conducted to assess a number of options at an early stage, effectively screening for potential methods to be taken into more detailed development testwork. The following beneficiation processes were examined: Screening and size based differentiation Scrubbing and attritioning Gravity beneficiation Magnetic separation Electrostatic separation Results of each stage of preliminary beneficiation testing are described below. Screening and de-sliming beneficiation Size by assay analysis was conducted on a 100 kg sub-sample from an equal mass obtained from all intervals within the downhole profile on two bulk samples derived from costeans excavated adjacent to drillholes LMAC0314 and LMAC0307. Table 18-13 below shows metal deportment by size for the sample grading 676 ppm U3O8. The following comments apply: Approximately 63.5% of the mass exists below a 300 µm cut size, with 55.3% occurring below 75 µm. Metal deportment typically concentrates into the fines, below 75 µm, with a substantial increase in the grade occurring in the size range 20 µm to 50 µm, possibly reflecting liberated carnotite. There is a steep reduction in the grade of slimes below 12 µm, representing 20% of the mass and – at a grade of 282 ppm U3O8 – slightly in excess of the mine cut-off grade at 200 ppm U3O8. On the basis that the overall head grade of this sample reported as 676 ppm U3O8, an equivalent -12 µm reject grade of approximately 209 ppm U3O8 could be expected, based on a 500 ppm U3O8 head grade. Given the possible impact of rheology on materials handling, water recovery, sodium carbonate recovery and RIP inter-tank screen throughput, as well as the relatively high silica content compared to coarser sizes, it is considered that early de-sliming and rejection of the slimes fraction may assist in improving overall circuit performance, relative to the incremental cost of accepting this fraction. Further rheology testing will be required to assess any

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improvement in rheology, assuming this fraction is rejected at an early stage in the process. Early rejection is also suggested, ahead of screening and comminution, given naturally occurring (possibly liberated carnotite) in the 20 µm to 50 µm size range. Table 18-13: Size x assay metal deportment Non-cumulative

Cumulative

Size (µm)

Mass (%)

U3O8 (ppm)

U3O8 distribution (%)

Mass (%)

U3O8 (ppm)

U3O8 distribution (%)

75

44.7

584

38.6

44.7

584

38.6

53

23.9

524

18.5

68.6

563

57.1

48

0.5

7,525

5.0

69.0

609

62.1

33

0.9

8,321

11.3

70.0

710

73.4

23

1.8

3,733

10.1

71.8

787

83.5

16

3.5

969

5.1

75.3

795

88.6

12

4.6

445

3.0

79.9

775

91.6

-12

20.1

282

8.4

100.0

676

100.0

Total

100.0

676

100.0

-

-

-

Electrostatic beneficiation Testing was conducted by drying the sample at 110°C and effecting a separation at 24,000 volts across a standard electrostatic plate separator. A concentrate mass of 4% was collected and assayed barren, excluding this method from further consideration. Magnetic beneficiation Monazite and some uranium minerals display slight magnetic properties when exposed to highly intensive magnetic fields. Given that the ore contains little iron mineralogy, it was viewed as a potential upgrade mechanism. However, Lake Maitland carnotite displayed little to no relative difference to the gangue mineralogy in magnetic susceptibility measured at 24,000 gauss, excluding this method from further consideration. Gravity beneficiation A Falcon concentrator was trialled on a 4.5 kg sub-sample with virtually no separation. Results yielded 9% uranium to concentrate in a 12% mass. Heavy liquid analysis showed that the specific gravity differences of carnotite at 2.80 SG and the bulk gangue at 2.65–2.70, coupled with particle size variations and inclusions, confirmed the inapplicability of high or low „G-force‟ gravity separation on whole ore. Colorimetric beneficiation Assay analysis of the 25 samples utilised in sighter leach testing showed that these samples displayed varying vanadium to uranium ratios. Examination of the physical appearance of +19 mm specimens from bulk samples showed a range of colours and six samples were examined to determine if colour could be used as an indicator to assess the vanadium ratio within carnotite. While colorimetric testing was not conducted at this stage, Table 18-14 below indicates higher vanadium to uranium ratio for darker samples.

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Table 18-14: Colorimetric beneficiation visual markers Visual identification

Activity (µSV/h)

V (ppm)

U (ppm)

Ratio V:U

Gypsum

0.1

72

< 20

> 3:1

Dolomite - Light

8.0

461

667

0.69:1

Dolomite - Light

5.8

528

815

0.66:1

Dolomite Intermediate

2.5

353

382

0.92:1

Dolomite - Dark

3.7

302

219

1.38:1

Dolomite - Dark

1.0

275

84

3.27:1

The results above indicate that colour does correlate to some extent with the composition of the carnotite. Given the relationship demonstrated between U:V ratio and leach extraction, further colorimetric testing may be applicable. Scrubbing and attritioning beneficiation A series of five attritioning and three scrubbing trials were conducted. Testing on lower competency samples indicated some potential for beneficiation and upgrade via these methods. However, attritioning tests conducted on competent calcrete fraction at various crush sizes down to 4 mm resulted in final reject assays generally exceeding 200 ppm U3O8. The potential for rejecting high grade calcrete (ferricrete and silicrete) due to its structural competence and poor surface abrasion qualities (shape) raised too high a risk to pursue the programme until more variability samples could be collected and tested. Scrubbing and Attritioning Upgrade - Competent and Soft Ore Fractions 100.0 90.0

% Metal Recovery

80.0 70.0 Scrubbed - Actual 60.0

Attritioned Competent Ore - Actual

50.0

Attritioned Soft Ore - Actual

40.0

Modelled Average Ore

30.0 20.0 10.0 0.0 30.0

40.0

50.0

60.0

70.0

80.0

90.0

100.0

% Yield

Figure 18-13: Scrubbing and attritioning metal recovery and yield

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18.3.5 Reverse sulphate flotation On the basis that previous sighter leach tests had indicated the presence of sulphates as the reason for elevated carbonate consumption and potentially lower leach rate (possibly due to celestine), a reverse sulphate flotation programme was conducted to evaluate the ability to beneficiate and remove sulphate mineralisation ahead of alkaline leaching. A total of six tests were conducted to define the physical parameters for best operation and it was concluded that flotation could be employed on the calcrete and clay fraction; however, in the case of clays (typically below 300 µm), flotation performance was negatively impacted by milling, requiring that the clay fraction bypass a milling stage. This stage of testing also indicated that the (-12 µm) slimes fraction would be required to bypass the flotation stage to maintain optimal performance. A total of four rougher tests were conducted on samples derived from costean LMAC0314 with the slimes (-12 µm) scalped out. The subsequent reverse flotation yielded: 86 – 92% sulphate removed to a high sulphate concentrate An average of 7% uranium loss to reverse sulphate concentrate A cleaner stage was tested on the combined high sulphate concentrate resulting in the recovery of 80% of the uranium from the rougher stage concentrate for recycle to the feed. Overall uranium loss from the reverse sulphate flotation stage was 1.4%, based on combined rougher and cleaner stages. Future work will investigate the potential of adding scavenging capacity to further reduce the sulphate grade in the flotation tail.

18.3.6 Ambient temperature pre-leaching Parallel to reverse sulphate flotation testwork and to mitigate the impact of sulphate, a series of tests were performed to investigate the ability to remove residual sulphate in the flotation tailing (alkaline leach feed) via an ambient temperature, reducing pre-leach. The concept in this case was to perform a pre-leach as a trim to remove residual sulphate via stoichiometric addition of sodium carbonate, maintaining slightly reducing redox conditions, followed by a CCD wash ahead of the elevated temperature alkaline leach stage. The chemistry of such a reaction was not fully understood at the time and so a generic regime was employed for this series of tests, allowing a baseline to be established for further optimisation. While „sacrificial‟ sodium carbonate was required in this case, the causticisation of carbon dioxide from power station off-gas was considered as a means to generate this stream. The technique employed was designed as a low temperature leach targeting a threshold below that required for typical uranium and vanadium dissolution but where sulphate could be removed to solution. Near stoichiometric reagent addition was employed to minimise reagent consumption. The trial consisted of eight tests and optimal conditions yielded a significant further sulphate (71%) reduction prior to the elevated temperature uranium leach. Uranium loss in this case was measured at 2-3% in solution which is ultimately lost from the process. The further extent of testwork on this unit operation is on hold pending results of further reverse sulphate flotation scavenging trials.

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18.3.7 Reverse flotation and elevated temperature alkaline leaching Initial process development testwork, reported here, targeted samples elevated in sulphate from costean LMAC0314 on the basis that previous sighter stage testing showed elevated sodium carbonate consumption rates as a result of direct consumption due to the presence of sulphate mineralisation. Elevated temperature alkaline leach testing was performed on samples from the LMAC0314 bulk sample previously treated to remove sulphates via reverse flotation (Section 18.3.6). Results of this suite of leach tests are presented below and show a significant improvement compared to previous results from sighter stage testing on whole ore samples not treated by reverse flotation. On the basis that the sample tested was derived from LMAC0314 with a higher average sulphate head grade, it is expected that variability testing will result in lower average sodium carbonate consumption rates than reported below. Table 18-15: LMAC0314 leach performance and reverse sulphate flotation Parameter

Unit

Whole ore leach

Reverse sulphate flotation and leach

P99 Size

µm

300

300

Sulphate head grade

%

4.43

0.59

Leach retention time

hour

32

32

Terminal sodium carbonate

g/l

34.4

43.4

Temperature

ºC

90

90

Sodium carbonate consumed

kg/t

92.9

32.7

Uranium extraction

%

70.0

97.9

A further optimisation programme was conducted to assess the impact of temperature, oxidant addition (hydrogen peroxide) and terminal carbonate concentration to optimise reagent consumption whilst maintaining leach performance. The results of this programme indicated that oxidative leaching via addition of hydrogen peroxide provided a substantial increase in leach kinetics. A residual slower leaching fraction was noted to be unaffected and was attributed to high vanadium ratio to uranium carnotite that may require a higher solution Eh (in excess of +150 mV vs AgCl) and increased oxidant addition. The results also indicated that a 95% uranium extraction target could be achieved within a nominal 24 hr leach residence time, relative to sighter stage whole ore leach tests on the sample from LMAC0314, where this target was not met after 32 hours retention.

18.3.8 Heap leach A preliminary assessment of heap leaching was conducted based on performing an intermittent bottle roll test on a sample of -19 mm/+355 µm derived from LMAC0314. The size fraction selected for testing was based on the approximate cut size where clay fraction could be excluded with testing proceeding on coarse (dolomite) fraction. Results at an average 33 g/L sodium carbonate concentration are summarised in Figure 18-14 and Table 18-16 below; the following comments apply: Cumulative uranium extraction reached 61.6%, while cumulative vanadium extraction reached 26.6% extraction after 48 days retention.

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While extraction was beginning to plateau after 48 days retention, cumulative sodium carbonate consumption rate increased to approximately 100 kg/t as a result of sulphates retained in the coarse fraction tested. On the basis of likely poor percolation characteristics with inclusion of clay and generally low metal extraction relative to reagent consumption, the test was terminated. Intermittent Bottle Roll Test LMAC0314 +355um 100.0

120.0 100.0

80.0 70.0

80.0

60.0 50.0

60.0

40.0 40.0

30.0 20.0

20.0

Sodium Carbonate (Kg/t)

Metal Extraction (%)

90.0

Uranium Extraction % Vanadium Extraction % Sodium Carbonate Kg/t

10.0 0.0 0.0

10.0

20.0

30.0

40.0

50.0

0.0 60.0

Retention (days)

Figure 18-14: Heap leach performance LMAC0314 +355µm

Table 18-16: Heap leach performance LMAC0314 +355µm Head grade (ppm)

Cumulative extraction and reagent consumption

Time (days)

U3O8

V2O5

U3O8 (%)

V2O5 (%)

Sodium carbonate (kg/t)

0

503

287

0.0

0.0

0

2

474

274

5.7

4.6

78

3

456

267

9.2

7.1

78

7

394

258

21.7

10.3

68

12

317

243

37.0

15.3

73

24

240

215

52.2

25.2

103

36

201

202

60.0

29.6

99

48

193

211

61.6

26.6

100

18.3.9 Slurry rheology Sighter stage mineralogy indicated a wide range of materials handling properties as a direct result of the wide variation in the proportion of clay and dolomite in samples examined. Rheology testing was conducted on samples from trench LMAC0314 prepared using slurry samples derived from: Fresh feed prepared below 300 µm Alkaline leach feed, prepared post reverse sulphate flotation Alkaline leach tailings

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Initial sample preparation involved settling and decanting to obtain slurry samples at 30% solids and 40% solids by weight. Preparation up to 50% solids was not possible by simple settling and decanting and a slurry sample at 50% solids was prepared by centrifuging. All samples were tested on a whole ore basis, without de-sliming, and further testing will be required to assess any improvement assuming early de-sliming and removal of -12 µm fraction. Results of forward scanned rheograms for samples prepared at 30% and 40% solids are presented below for each separate slurry type.

Shear Stress (Pa)

Rheogram Fresh Feed Slurry 160.00 140.00 120.00 100.00 80.00 60.00 40.00 20.00 0.00

Actual 40% Solids Actual 30% Solids Estimated 20% Solids Estimated 25% Solids

0.0

200.0

400.0

600.0

800.0

Shear Rate s^-1 Figure 18-15: Slurry rheology fresh feed -300µm

Testing at ambient temperature on fresh feed showed that the slurry was shear thinning and followed a Herschel-Bulkley correlation. An approximate limit of 30% solids is indicated for application of centrifugal pumping and a nominal limit between 30% and 40% solids is indicated as a maximum underflow density achievable by thickening alone.

Shear Stress (Pa)

Rheogram Alkaline Leach Feed 160.00 140.00 120.00 100.00 80.00 60.00 40.00 20.00 0.00 0.00

Actual 40% Solids Actual 30% Solids

200.00

400.00

600.00

800.00

Shear Rate s^-1 Figure 18-16: Slurry rheology alkaline leach feed

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Rheogram Alkaline Leach Tailings

Shear Stress (Pa)

300.00 250.00

Actual 40% Solids

200.00

Actual 30% Solids

150.00 100.00 50.00 0.00 0.00

200.00

400.00

600.00

800.00

Shear Rate s^-1 Figure 18-17: Slurry rheology alkaline leach tailings

Results of initial slurry rheology testing indicate challenging conditions for pumping and thickening unit operations. Rheograms progressively track slurry characteristics from fresh feed, through sulphate flotation and elevated temperature alkaline leach stages. The data presented indicate that the slurry progressively becomes more viscose with processing; this is seen at this stage as a result of shearing the slurry and increased viscosity due to the addition of sodium carbonate in the leach. The data indicate that filtration would not be applicable for the slurries tested and further testwork has been planned to assess alternative methods for dewatering to be assessed in conjunction with thickening and CCD unit operations. The use of densifying cyclones and centrifuging on fines fractions is currently under consideration. This stage of testing will be performed utilising a pilot scale thickener to assess settling rate, ability to achieve compression and ultimate underflow density. Testing utilising other unit operations in conjunction with CCD will also be assessed against de-sliming -12 µm fraction to measure any improvement in rheology as a result of early removal of the ultra-fines fraction.

18.3.10 Uranium refining and recovery Uranium refining process selection was considered based on a desktop study undertaken to compare a CCD/IX circuit compared to final uranium recovery via an RIP system. On the basis that refinery selection is driven by the potential for soluble uranium loss and this loss is controlled in a leach process by slurry rheology, the study utilised results of initial slurry rheology testing reported in Section 18.3.8. CCD/IX and RIP refining options indicated an RIP system provided a reduced risk in terms of uranium loss and reduced metal recovery and increased the requirement for additional process water required for the CCD/IX option. On this basis, an RIP refining process was selected for further assessment.

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18.3.11 Flowsheet testing Combined flowsheet tests were conducted to assess and compare overall metal extraction based on attritioning, flotation and leaching against alternative de-sliming, splitting and milling of calcrete fraction, flotation and leaching. PLS generated in each test was provided for initial refinery modelling and development of a SYSCaD model of the overall process flowsheet. The following flowsheet tests were conducted: Attritioning beneficiation of coarse calcrete, followed by reverse sulphate flotation and leaching resulting in a process recovery of 87.0% with a 97.9% leach extraction. Whole ore treatment involving de-sliming, calcrete milling, flotation and leaching. This resulted in a process recovery of 95.5% and a leach extraction of 95.6%. A further trial needs to be conducted assessing de-slimed material rejection.

18.4

Process flowsheet

18.4.1 Process flow Results of early flowsheet testing indicate improved performance via initial de-sliming at 12 µm, milling of +300 µm calcrete and reverse sulphate flotation to remove sulphate ahead of elevated temperature alkaline leaching. On completing the scoping testwork programme, a series of process flow diagrams and a SYSCaD process model were developed for economic analysis and engineering assessment. The block flow diagram below (Figure 18-18) represents the main process flows considered to date.

Figure 18-18: Proposed Lake Maitland block flow diagram

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18.4.2 Process description Ore preparation ROM ore is received by the ore preparation circuit where a natural low grade clay fraction is removed and the remaining ore is converted into a slurry. Reverse sulphate flotation Sulphate impurity minerals are removed via froth flotation. Sulphate minerals consume sodium carbonate leaching agent if not removed. Chloride washing Chloride salts which occur naturally in the ore are removed by a counter current decantation wash circuit. Low chloride water is contacted counter with the flow of ore solids through the circuit. Chlorides need to be removed from the leach circuit as they reduce the efficiency of the downstream recovery circuit. Alkaline leaching The washed ore slurry is leached at 95˚C with sodium carbonate in a series of stirred tanks. This leaches over 95% of the uranium and approximately 60% of the vanadium into solution. Resin in pulp adsorption Leached slurry is contacted counter currently with ion exchange resin beads in a series of tanks. The uranium adsorbs onto the resin leaving many of the impurities behind in the tailings slurry. The RIP adsorption circuit recovers greater than 99% of the soluble uranium in the uranium leach slurry. Resin elution Resin beads are separated from the slurry by screening and washing with water. Uranium is eluted from the resin by contacting with concentrated sodium bicarbonate solution. This creates a concentrated uranium product solution, with the stripped ion exchange resin returning to the RIP adsorption circuit for further uranium loading. UO4 (uranium peroxide) precipitation To separate impurities contained in the resin eluate, uranium is precipitated as sodium diuranate (SDU) from the concentrated uranium stream via the dosing of sodium hydroxide. The SDU is separated then dissolved in sulphuric acid to remove residual solids. Hydrogen peroxide is then dosed into the uranium containing sulphuric acid to precipitate high purity uranium peroxide crystals or UO4.xH2O. UO4 product handling The uranium peroxide crystals are removed from the acid solution by thickening and filtration. The final uranium peroxide product is then dried at 100˚C to the required moisture content and packaged in sealed product drums for export.

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Reagent recovery Sodium carbonate is recycled back to the leach stage by separating carbonate rich liquor from the RIP barren slurry. Tailings treatment The solids from the reagent recovery are combined with other tailings streams to form slurry ready for pumping to a tailings storage facility.

18.4.3 Conclusions Based on the scoping testwork reported here, the following conclusions are offered: Ore handling and comminution The nature of the ore, containing significant and highly variable proportions of clays, will require careful consideration of ore handling and repulping. Scrubbing has proven to be a useful means to repulp and effect a split between dolomite and clays and should be included for this purpose. The ability to improve overall front end ore handling as well as providing an improvement in overall circuit slurry rheology may be possible by early de-sliming and rejection of the -12 µm slimes fraction. Examination of the distribution of sulphate in sighter testing indicated some potential to separately scrub and scalp gypsum concentrated at the surface. Testing to date has been performed on samples containing substantial gypsum fractions (costean LMAC0314) and any ability to scalp gypsum would assist in reducing overall carbonate consumption rate, over and above that achieved to date via reverse sulphate flotation. Limited comminution testing has indicated a low power requirement and very low abrasion potential. This implies that toothed roll crushers or sizers could be used in place of jaw crushing. Based on high reject grades obtained for competent calcrete via scrubbing and attritioning as well as a requirement to perform reverse flotation on the natural -300 µm fraction, milling preparation of coarse calcrete fraction is preferred. Beneficiation Beneficiation testing has eliminated a number of potential methods, including gravity separation. Observations based on coarse fraction colour and grade indicated that colorimetric testing may provide a coarse method for distinguishing high and lower grade fractions. Scrubbing and attritioning do act to upgrade the feed, although testing on competent fractions indicated that the resultant reject grade was in excess of 200 ppm U3O8, even with size reduction down to 4 mm. This methodology may be applicable to softer fractions; however, the risk of rejecting metal derived from competent fractions increases the risk of metal loss for no complementary improvement in other cost drivers. Reverse flotation Due to the negative impact that comminution of the clay fraction has on flotation, flotation should be performed by splitting the natural -300 µm fraction direct to flotation with milling preparation of coarse calcrete. Reverse sulphate flotation testing to date is not optimised; however, this approach has shown the potential to reduce leach sodium carbonate consumption to approximately a third of that

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required for whole ore processing. This methodology provides a robust treatment and buffer to control the consumption of sodium carbonate lixiviant in the leach stage. Additional control over sodium carbonate loss to tailings is required and predicated on slurry rheology and ability to mitigate impacts. Leaching Comparison between heap leaching, scrubbing + attritioning + agitated leaching and reverse sulphate flotation + agitated leaching shows that reverse flotation + agitated leaching provides a higher overall metal recovery, reduced operating cost and is preferred. Leach testwork was conducted on a sample derived from costean LMAC0314 and represented the worst case in terms of sulphate content tested to date. Comparison between leach extractions achieved with and without flotation shows a substantial improvement, likely as a result of removal of celestine in conjunction with a more oxidising leach. Leach sodium carbonate consumption is controlled by the level of total sulphate in the feed. While gypsum appears to impact carbonate consumption, celestine appears to impact both leach rate and carbonate consumption rate. Based on the fact that flotation testing was undertaken on samples derived from costean LMAC0314 and represented the worst case in terms of sulphate content relative to other areas tested during sighter leach testing, it is expected that lower sodium carbonate consumption rates, below 32 kg/t should be possible when treating the resource average sulphate grade. Refining Cost benefit analysis examining CCD/IX and RIP technologies has shown that RIP is preferred, largely due to clays and resultant rheology. Slurry rheology and ability to deslime may impact on RIP inter-tank screen flow capacity.

18.4.4 Recommendations Based on results of scoping testwork performed to date, the following recommendations are made with regard to further testwork required to complete and validate the proposed flowsheet: Variability testing Testing to date has concentrated on samples derived from the northern and southern areas of the resource. Definitive flowsheet testing and validation are now required to be undertaken on representative samples from all the major domains within the resource. Ore handling and comminution A decision is required regarding the practicality of separate gypsum scalping. In the event that scrubbing and scalping is practical, then this should be included in further flowsheet testing to assist in minimising sodium carbonate consumption rate. Additional comminution testing is required on a variability basis and will need to include: JK SMC, UCS and full Bond suit. Additional abrasion index testing will be required to validate the very low value obtained to date to confirm the applicability of sizers for front end preparation. Bulk materials handling testing is required to provide design criteria for materials selection, i.e. friction angle and corrosion potential as well as design criteria for bin and chute design. Testing typically provided by TUNRA is recommended in this case, based on de-slimed feed fraction.

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Milling preparation of calcrete will require some optimisation to ensure optimal liberation of carnotite and celestine, ahead of reverse sulphate flotation. Flotation Additional testing is recommended to assess the scavenger stage flotation combined with rougher and cleaner stages utilised to date. An assessment of the use of Jamison cells is recommended for cleaner stage duty. Optimisation (matrix) testing is required to provide optimal flotation and final design criteria. Optimisation will need to be performed ahead of variability testing and will also need to allow for possible de-sliming and/or gypsum scalping. Comprehensive variability testing on a flowsheet basis should now be undertaken on representative material from all the major domains within the resource. Leaching Optimisation (matrix) testing is required to provide final design criteria. Test variables should include: pH, Eh, calcrete grind size, temperature, carbonate tenor and bi-carbonate ratio. Optimisation will need to be performed ahead of variability testing and will also need to allow for possible de-sliming and/or gypsum scalping. Comprehensive variability testing on a flowsheet basis should now be undertaken on representative material from all the major domains within the resource. Pilot scale thickening and densifying cyclone testwork is required to provide firm data describing mass and water split and maximum underflow density to assess overall water balance requirements and ability to maximise sodium carbonate washing and recovery (post leach). Pilot scale centrifuging testwork is required to assess the ability to achieve high densities on fines fraction to improve tailings recovery of sodium carbonate. Refining RIP testwork and optimisation are required to select resin type and provide firm design criteria. Inter tank screen flow capacity testing is required to assess the impact of slurry rheology and any improvement likely via early de-sliming of -12µm fraction. Vanadium precipitation testwork is required to provide design criteria for this unit operation. Tailings disposal Flowsheet testing is required to provide bulk samples of tailings for environmental and geotechnical testing and provision of tailings design criteria.

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Page 94

Mineral Resource and Mineral Reserve Estimates

In 2009 SRK was requested by Mega to update the estimate of Mineral Resources for the LMUP in Western Australia. The following description of the resource estimate is adapted from Gleeson (2009).

19.1

Geological modelling and domaining

The mineralisation at LMUP is highly dependent on the lithofacies and hence a new approach was adopted to attempt to appropriately domain the drilling data. While traditional methods of wireframing such a complex sequence of lithofacies had previously been used, these proved too time consuming and difficult. In this resource, geological domaining is necessary as each lithology has its own unique uranium grade distribution and characteristics. Due to the difficulty of providing coherent wireframe models of the geology, previous estimates only made attempts at 2D global estimation – thus reducing the confidence levels in the estimates on a local basis.

19.1.1 Geology modelling Two processes were used in building a 3D lithofacies model. Firstly the base of the channel in which the resource is constrained was modelled using basements intercepts from drilling (Figure 19-1). Implicit geological modelling package Geomodeller was used. With over 1,400 drillholes, using a mathematical modelling technique sped up the time to complete the model. Further details on the Geomodeller technology can be found in Chilès et al. (2004). Rather than build wireframes (3D triangulations) of each lithofacies, a more appropriate method is to model them using a 3D volume model (block model). Firstly, an empty 3D model was built using the base of channel wireframe. This model differs from traditional block models in that it is a stratigraphic model (S-grid2) in which the cell dimensions and orientations are allowed to vary. This allows continuity along stratigraphic horizons to be maintained. In the case of the LMUP, this effectively allowed the modelled sediments to „drape‟ into the basement (Figure 19-2). The nominal cell dimensions used were 85m by 65m by 0.25m (NS x EW x Vertical). Secondly, the different lithofacies were estimated into the model using a categorical modelling tool. Thirteen different lithofacies (Table 19-1) were estimated using an inverse distance algorithm based on indicator technology (Figure 19-2). A visual comparison of model geology and drilling on a series of sections showed an excellent correlation. Table 19-1: Lithofacies estimated into S-Grid Calcrete

Evaporite – Gypsum

Gritty Clay

Sand

Black Shale

Silcrete

Clay

Sandy Clay – Clayey Sand

Ferricrete

Mudstone

Kopi

Soil

Calcrete clay

2

The GoCAD software package was used in modelling the LMUP. S-grids are a technology employed in GoCAD and are widely used in the Petroleum industry.

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19.1.2 Grade domaining As there are mineralised and un-mineralised areas within the different lithofacies, a further domain based on eU3O8 grade was applied to the 3D model. Using 0.25 m composited downhole gamma eU3O8 data, a 100 ppm eU3O8 grade shell was built using the LeapfrogTM software package. This shell was constrained within the channel model built in Geomodeller. LeapfrogTM is an implicit (mathematical) grade modelling tool which was used to quickly generate the required grade shell (Cowan et al., 2003, describes the technology used in LeapfrogTM). Given the very thin geometry of the deposit, the modelling parameters were highly isotropic. The 100 ppm eU3O8 grade shell provided a good level of mineralisation continuity. The 3D volume model and grade shell were combined to produce a 3D volume model containing separate regions for each lithofacies inside the 100 ppm eU3O8 grade shell (Figure 19-3). The final domains used in the resource estimate are a combination of the geological and grade domain model.

Figure 19-1: Plan and 3D view of the base of channel surface constructed in Geomodeller Note: 60 x vertical exaggeration.

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Figure 19-2: 3D view looking NE along the channel showing the categorical modelling of rock types in the channel Notes: 60 x vertical exaggeration; block colour coded by geology.

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Figure 19-3: 3D view of the combined GoCAD categorical geology model constrained by LeapfrogTM generated 100 ppm grade shell Notes: 60 x vertical exaggeration; block colour coded by geology.

19.2

Estimation

19.2.1 Database All processed radiometric data were provided in a number of Microsoft Access databases by consulting Geophysicist David Wilson of 3D Exploration (Section 15). All remaining drillhole data were provided by Mega as a single Microsoft Access database. Overall the database management practices observed are poor, but improving. SRK would recommend that a single database be created with different tables for repeats, duplicates, standards, chemical, gamma logs, etc. Only gamma log measurements from Redport and Mega were used in the resource estimation.

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19.2.2 Densities Two sets of density data were reviewed during the course of the resource estimation: 1.

Calcrete densities taken from a series of 13 historical trenches and test pits around the resource in March 2009 – a total of 67 density measurements.

2.

Whole core densities from a variety of different rock types in 17 different holes throughout the resource – a total of 149 density measurements.

The improvements in the quality and quantity of bulk density data determinations available for most rock types have enabled SRK to estimate tonnages for the majority of the geological domains and rock types within the resource to a moderate-high level of confidence. This has led to over 90% of the resource having reliable density estimation. Calcrete densities In March 2009, sampling and geological logging of calcrete were undertaken at 13 historical Carpentaria Exploration Corporation (CEC) costeans located within the Lake Maitland deposit, on exploration licence E53/1099. The purpose of the programme was to collect sufficient bulk samples of calcrete from multiple localities to provide statistically meaningful data on the physical nature and variability of the samples. Dry bulk density measurements were undertaken at AMMTEC laboratories. The costeans were identified from historical CEC maps and with reference to nearby Mega and Report drilling. Location on the ground was done by handheld GPS. Trench locations are shown in Figure 19-4. Where possible, six pieces of calcrete were selected from the bulk samples taken from each trench. To reduce measurement error, larger pieces were taken, and selection of samples was made to ensure variation in the physical types of calcrete was represented. Overall, 67 calcrete samples were collected for a combined weight of 99.37 kg. Each selected calcrete sample was firstly dry-scrubbed to remove any loose and excess surficial material, then geologically logged (calcrete type, induration, colour, carnotite mineralisation), weighed and photographed. All the samples were inserted into calico bags numbered with a CEC prefix relating to the costean location, followed by a unique sample ID number. All samples were taken from near surface, unconsolidated, backfilled material. Calcrete samples generally were of a competent rather than friable nature, mainly due to varying amounts of cementation. There were often hardness variations within each sample reflecting bands or pockets of increased carbonate cementation. Carnotite mineralisation was observed in most of the samples, occurring mainly as „smears‟ on the planar surfaces of the calcrete. The results from AMMTEC Laboratories for bulk dry density measurements are presented in Appendix 3.

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Figure 19-4: CEC trench locations sampled by Mega for calcrete density testwork Notes: Red outline is current 100 ppm eU3O8 cut-off Resource Estimate.

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Sonic core densities The sonic core bulk density data are mostly collected at 0.5 m increments. Because of the low number of different sample lengths, all density data have been used in this analysis with no weighting for length applied. Table 19-2 presents the summary statistics by logged lithological code. The results clearly indicate that of the seven rock codes, the following five discrete lithological groupings for density can be defined: 1. 2. 3. 4. 5.

Clays – comprised of codes CLY and CCLY for a total of 62 data Sands – comprised of codes SACY and CYSA for a total of 70 data Calcrete – code CLCR, with 9 data Conglomerate – code CONGL, with 1 datum Evaporite – code EVPT, with 1 datum Table 19-2: Sonic core density summary statistics by lithology Lithology

Count

Min

Max

Mean

SD

Var

CV

Median

CLY

45

0.62

2.24

1.33

0.32

0.1

0.24

1.29

CCLY

17

0.88

1.81

1.37

0.28

0.08

0.21

1.30

SACY

56

0.83

2.31

1.69

0.34

0.11

0.2

1.77

CYSA

14

1.25

2.5

1.85

0.34

0.11

0.18

1.87

CLCR

9

1.25

1.93

1.61

0.26

0.07

0.16

1.75

EVPT

1

1.18

1.18

1.18

0

0

0

1.18

CONGL

1

2.15

2.15

2.15

0

0

0

2.15

The statistics for the five groupings are presented in Table 19-3. Table 19-3: Density summary statistics by lithological grouping Lithology

Count

Min

Max

Mean

SD

Var

CV

Median

Clays

62

0.62

2.24

1.34

0.31

0.1

0.23

1.30

Sands

70

0.83

2.5

1.72

0.34

0.12

0.2

1.79

Calcrete

9

1.25

1.93

1.61

0.26

0.07

0.16

1.75

Evaporite

1

1.18

1.18

1.18

0

0

0

1.18

Conglomerate

1

2.15

2.15

2.15

0

0

0

2.15

Density data assignment The mean value for each rock grouping, as presented above in Table 19-3, has been used to assign a constant density value to all blocks of the same rock type. In doing so, it is worth noting that: 1.

For the clay and sand domains there is a reasonable amount of data spread throughout the deposit. None the less, each domain has limited data restricting the classification to a maximum of Indicated Resources.

2.

For the calcrete domain, only nine density data are available from the core data. However, a further 67 samples are available from the CEC trench samples.

3.

Both the conglomerate and evaporate domains have only 1 datum each.

In terms of density, SRK recommends that more density data are collected in future drilling and trenching programmes to improve the reliability of tonnages assigned to the resource. Targeting density collection of under-sampled or unsampled rock types should also be carried out. However, SRK believes there are now sufficient density data available for most rock types to be used in a HERO/GLEE/GUIB/WILL/mool

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meaningful way to estimate tonnages reliably for most of the principal geological domains and lithotypes. This has had an important bearing on the reclassification of significant parts of the resource from Inferred to Indicated. The densities used in the resource estimate are shown in Table 19-4. Table 19-4: Density used by rock type Domain rock type

Density (t/m3)

Calcrete

2.18

Clayey calcrete

1.61

Clay

1.33

Evaporite gypsum

1.33

Clayey sand/sandy clay

1.77

Conglomerate

2.15

Sand

1.72

Mudstone

3

1.76

3

1.76

Gritty clay

3

Black shale

1.76

Kopi3

1.76 3

Ferricrete

1.76

3

Silcrete

1.76 3

Carbonate rock

1.76

3

Soil

1.76

19.2.3 Disequilibrium Some 46 channel samples of material derived from trenches at the Lake Maitland resource were submitted to CSIRO in May 1980 for the purpose of establishing the presence of secular disequilibrium. Disequilibrium was determined by comparing the U3O8 value determined by DFN (Delayed Fission Neutron) analysis with the equivalent (eU3O8) uranium values obtained from sealed can gamma analysis (Figure 19-5). The results showed that the low grade samples (less than 200 ppm U3O8) are relatively radium rich, whereas the higher grade samples are closer to radioactive equilibrium. The mean disequilibrium ratio derived from the tests for samples >200 ppm was 0.97.

3

3

For litho-facies with no density testing, a value of 1.76 t/m was used.

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U3O8 (ppm)

Figure 19-5: Scatterplot showing gamma eU3O8 values plotted against DFN U3O8 values

Therefore, on average the samples are only 3% radium rich. This result indicates that disequilibrium does not appear to be a significant problem in the resource and that gamma logging with applied calibration factors should give a reliable estimate of contained U3O8. The routine chemical sampling of drillholes should continue, in addition to gamma logging, to ensure there are no zones within the resource where this may occur. There is no evidence to suggest, from comparing chemical assays with the gamma log assays, that significant disequilibrium is an issue. Whilst absolute values for U3O8 may vary between the gamma logs and the chemical samples over any given interval, it is clear that the intervals do show the same general levels of U3O8 and in the same vertical location. To date no additional disequilibrium work has been undertaken on the latest Mega drilling. It is strongly recommended that some new disequilibrium studies be carried out as a routine check on at least 50 or so of the drill samples taken in the last three years from a variety of different geographic locations and rock types within the resource. The sonic core provides a superior sample volume to the aircore drill samples and should provide a definitive indication of the presence of disequilibrium both vertically and laterally throughout the resource.

19.2.4 Statistical analysis of resource domains This section covers a statistical review of the gamma log assay data for all the geological domains identified and used in the resource estimate (i.e. the categorically modelled lithofacies). The 100 ppm eU3O8 LeapfrogTM envelope is not taken into account. De-surveying and compositing of the data files resulted in a drillhole database comprising 58,000, 0.25 m downhole composites for eU3O8. The 0.25 m composite sample lengths were chosen as reflecting the possible minimum mining width that could be applied to the resource. The total drilling metreage is 19,541 m from 1,491 holes. Holes were generally short – 7 to 15 m – with the

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deepest being 120 m. Most were air core, with the remainder being sonic core. All grade values used were from the downhole gamma logging data. The summary statistics for the major domains within the channel are given in Table 19-5. Table 19-5: Statistics for 0.25 m composites by geological domain Domain

Min value

Max value

Arith mean

SD

Variance

CV

Number of samples

All domains

0

6,362

75

203

41,239

2.7

43,370

Calcrete

0

3,988

156

349

122,248

2.2

6,777

Clay

0

6,362

95

247

61,439

2.6

15,594

Clayey Sand, Sandy Clay

1

2,391

43

102

10,562

2.4

14,621

Calcrete Clay

1

2,108

146

249

62,447

1.7

570

Figure 19-6 through to Figure 19-10 show the frequency histograms and probability plots for these domains. All the domains show a positive skew in the data distribution. A review of the domains shows clearly that calcrete is the domain with the highest average grades, followed by the clayey calcrete and clay units. In general, the gravel, grits and elluvium display the lowest grades. A review of the individual histograms shows that there are few samples above 2,000 ppm eU3O8. However, a few extreme high grade outliers between 2,000 and 6,000 ppm eU3O8 do occur and are mainly associated with the calcrete and clay zones. It is also clear from the statistical analysis of individual domains that the domains have quite different distributions in terms of average grade, variability and CV to warrant individual treatment from an estimation point of view. The CV – as a measure of the relative variability of each domain – shows quite a range from lows of 0.5 up to 3.6. Usually deposits with a CV greater than 1 can have problems with local estimation due to relatively high nugget effects. In virtually all of the main resource domains at Lake Maitland the CV exceeds 1.5 to 2. Therefore, some problems with local estimation of grade on a block-by-block basis can be expected. The number of samples in most of the domains is sufficient to ensure that a statistically reliable measure of grade and statistical variability can be made. The only exceptions are the silcrete and elluvium domains, which have less than 100 samples.

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eU3O8

eU3O8 eU3O8

eU3O8

Figure 19-6: Frequency histogram and log probability plot for all domains in the general Channel domain

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eU3O8

eU3O8

eU3O8

Figure 19-7: Frequency histogram and log probability plot for Calcrete domain

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eU3O8

eU3O8 eU3O8

eU3O8

Figure 19-8: Frequency histogram and log probability plot for Calcrete Clay domain

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eU3O8

eU3O8

eU3O8

Figure 19-9: Frequency histogram and log probability plot for Clay domain

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eU3O8

eU3O8

eU3O8

Figure 19-10: Frequency histogram and log probability plot for Sandy Clay/Clayey Sand domain

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19.2.5 Application of upper cuts No upper cut has been applied to the eU3O8 downhole composites used in the estimation. This has been justified on the following basis: No isolated high grade values exist that would severely impact on having a large area of influence, resulting in areas of large tonnages and high grade. Most high grade values occur adjacent to lower grade intercepts, thus reducing the potential for high grading. A review of the spatial distribution of the high grade samples (+100 ppm eU3O8) shows little spatial clustering and most are distributed more or less evenly throughout the resource. These appear to be quite real and statistically numerous to indicate that a high grade component exists within the resource. Over 780, 0.25 m samples above 1,000 ppm eU3O8 occur in the resource and are distributed relatively evenly within a central high grade core that extends over much of the central portion of the resource. A smoothing algorithm (Kriging) is being used as an estimator on a block of a size that is approximately one half of the average sample spacing. Each block estimate is based on at least 15 to 20 samples, which ensures that the effect of any individual high grades is reduced. There appear to be no extreme high grade outliers in the data set. This may be due to the gamma logs being less susceptible to high grade nugget effect than chemical samples of drill cuttings as the downhole gamma eU3O8 values effectively sample a much larger volume (approximately 50 x that of a similar length of aircore sample); hence the variability of the gamma log samples is far less than that observed in destructive sampling over the same interval.

19.2.6 Variography Directional variograms were trialled for all geological domains and also the resource as a global entity, using the 0.25 m eU3O8 composites. Downhole variograms were trialled in addition to directional variograms. For some domains, it was difficult to obtain meaningful variography due to lack of samples available in those domains. The lack of close-spaced samples in the E–W and N–S direction made the interpretation of short range structures difficult in these directions. The minimum hole spacing in these directions was only 25 m, so structures shorter than this were hard to identify. The best short range structures identified were in the vertical downhole direction where samples were available on a 0.25 m interval. Where good directional variography could not be obtained for a domain, the global variography as defined for the entire resource was used. For all the major domains (calcrete, clay, sandy-clay, calcrete-clay, etc) sufficient data were available to determine unique variography. In terms of tonnage, it was only necessary to apply the global variogram models in some of the minor zones. Directional variograms were developed with the main axis of continuity being slightly more in the N–S direction (along channel) than E–W (across channel). In general, the long range structures for these two principal horizontal directions were in the order of more than 1,000 m, whilst mid-range structures were of the order of 500 m and the short range structures for the vertical direction were in the order of a few metres, rarely exceeding 3 or 4 m. These observations fit well with the overall observed continuity of mineralisation observed from the drilling. Directional variograms were developed for all domains with a search tolerance of 22.5 o in the horizontal and vertical directions. A variety of different lags were trialled to obtain the best structures. In general, a lag distance of 90 m worked best for the horizontal direction, whilst 0.25 m worked best for the vertical direction (as would be expected from the average sample spacing).

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Figure 19-11 through to Figure 19-14 show the experimental and theoretical fitted variograms for each of the principal directions of anisotropy (vertical, N–S and E–W) and for the major domains.

Figure 19-11: Directional variography (actual and fitted theoretical) for the Global Channel domain

Figure 19-12: Directional variography (actual and fitted theoretical) for the Calcrete domain

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Figure 19-13: Directional variography (actual and fitted theoretical) for the Clay domain

Figure 19-14: Directional variography (actual and fitted theoretical) for the Sandy Clay/Clayey Sand domain

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All the theoretical and actual variograms developed were of the normal form. In some cases, pairwise relative variograms (not shown in this report) were trialled and gave the best structure, especially along strike. In these cases, only the ranges were derived from the pairwise relative variograms. The actual nugget and sill variances were taken from the normal variograms, often seen in the vertical direction. The theoretical fitted variogram parameters for each domain can be seen below. The theoretical curves fitted were all of a three structure spherical form. Note that not all variograms gave well-structured results and in some cases the best visual fit is – to some degree – subject to personal interpretation. Some of the domains displayed a strong zonal anisotropy. In these cases it was often necessary to extend the long range structure ranges out a great distance in the horizontal direction to match the sill seen in the vertical well-formed variography. The variogram parameters for the main domains are shown in Table 19-6.

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Table 19-6: Variogram parameters Global Channel domain Co=10000 Structure 1

R1 (90)

R2 (180)

R3 (vertical)

Ranges

101

121

1.5

Variance

28261

Structure 2

R1 (90)

R2 (180)

R3 (vertical)

Ranges

503

503

1.57

Variance

45053

Structure 3

R1 (90)

R2 (180)

R3 (vertical)

Ranges

1475

1475

1.58

Variance

6300 Calcrete domain

Co=10000 Structure 1

R1 (90)

R2 (180)

R3 (vertical)

Ranges

172

105

0.3

Variance

35193

Structure 2

R1 (90)

R2 (180)

R3 (vertical)

Ranges

495

495

1.3

Variance

84841

Structure 3

R1 (90)

R2 (180)

R3 (vertical)

Ranges

1890

1612

4.2

Variance

170940 Clay domain

Co=10000 Structure 1

R1 (90)

R2 (180)

R3 (vertical)

Ranges

183

145

0.9

Variance

35193

Structure 2

R1 (90)

R2 (180)

R3 (vertical)

Ranges

809

535

2.8

Variance

40138

Structure 3

R1 (90)

R2 (180)

R3 (vertical)

Ranges

7599

4877

7.7

Variance

150000 Sandy Clay/Clayey Sand domain

Co=1000 Structure 1

R1 (90)

R2 (180)

R3 (vertical)

33

1.2

Ranges

63

Variance

5394

Structure 2

R1 (90)

R2 (180)

R3 (vertical)

Variance

219

200

6.6

Structure 3

R1 (90)

R2(180)

R3 (vertical)

Ranges

1711

1210

8.0

Variance

13800

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19.2.7 Block modelling As described in a previous section, 13 domains based on the geological modelling combined with 100 ppm eU3O8 LeapfrogTM wireframes were used in the estimation. In addition, a domain termed „channel‟ was used for a few lithologies that represented less than 1% of the total resource volume. A stratigraphic block model was developed in GoCAD with the following nominal block dimensions: East (X) 65 m North (Y) 85 m Vertical (Z) 0.25 m. The choice of block size was based on being approximately one half of the average sample (drillhole) spacing in the X and Y direction and equivalent to the sample composite sample length in the vertical direction. The vertical direction also corresponds to a suggested minimum mining height for the resource. More generally, this model assumes that the mining selectivity will not be better in the vertical than the block height. The block sizes are large enough to ensure that an unbiased grade estimate is achieved. The model limits in relation to the drill collar locations are shown in Figure 19-15.

Figure 19-15: Plan view of resource block model grid with drillhole collar locations overlain

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A GoCAD S-Grid (stratigraphic) grid was chosen over a regular grid as this type of grid has the ability for an estimator to follow a stratigraphic horizon, which is preferable in a stratiform resource such as this. The stratigraphic grid maintains correct stratigraphical continuity of grade within the mineralised horizons/zone. The grid is designed to be bedding plane parallel. This type of grid is commonly used in petroleum reservoir modelling where it is often essential to perform estimates along stratigraphic horizons. S-Gridding deforms the block shapes to follow the stratigraphy. It therefore results in minor block volume changes for some blocks. A review of the different block volumes within the resource does show some volume variation of no more than 5% between blocks, with 90% of all blocks not having any significant variation on volume. For the purposes of adhering to regular volumes and not having significant volume variance effect, the S-Grid is considered more than adequate compared to a regular grid and is therefore considered suitable for use in Kriging estimates. However the S-Grid is finally regularised post-estimation for export and use in mining studies in regular mining software. Figure 19-16 shows a plan and 3D view of the geology domains within the S-Grid.

Figure 19-16: Plan and 3D view of the S-Grid block model domained by geology

19.2.8 Estimation/Kriging The resource was estimated using ordinary Kriging. This linear unbiased estimator was considered suitable for estimating eU3O8 grade due to the relatively low CV encountered in each domain suggesting few estimation problems due to very high variability, high grade outliers or severe nugget effect. In general most domains showed the nugget variance to be less than 25% of the total variance. The model was estimated on a domain by domain basis using the 13 previously defined domains and the general „channel‟ domain. All the 0.25 m sample composites used for the estimate were flagged by domain and estimated directly into the relevant domain using the appropriate variogram established for each domain. All estimates were further constrained within the 100 ppm eU3O8 wireframe shell. Only eU3O8 was estimated. All densities were applied directly to the domain as a single value per domain, based on the unique density measurements for each particular domain. HERO/GLEE/GUIB/WILL/mool

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The Kriging parameters used were: Estimation method: Ordinary Kriging Search parameters: Search ellipse radius (150 m X, 150 m Y, 5 m Z) Minimum sample data used per block estimate: 4 Maximum sample data per block estimate: 32 Maximum per octant: 4 Estimation was restricted within the 100 ppm eU3O8 grade shell and also by distance. No cells were estimated beyond a range of 150 m in the horizontal direction and 5 m in the vertical direction. In addition, estimation follows the stratigraphic layers within the S-Grid to give a more geologically appropriate estimate. No upper cutting of grade was applied to any of the domains as no extreme high grade outliers were identified. Figure 19-16 shows the plan and 3D view of the Kriged estimates for eU3O8 on the channel block model. It is clear from a review of the block model, that the estimate is highly constrained geologically within the channel and 100 ppm eU3O8 grade shell.

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Figure 19-17: Plan and 3D view of the Kriged estimates for eU3O8 on the channel block model

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Classification

Since the 2007 resource estimate by Hellman and Schofield which classified the entire resource as Inferred, it may be possible in SRK‟s opinion, and in keeping with the NI 43-101 and JORC codes for reporting mineral resources and reserves that much of the resource can now be upgraded from Inferred to Indicated category. In the current estimate, some 90% of the resource is now placed in the Indicated category and only 10% remains in the Inferred category. The reasoning behind the decision to upgrade the resource estimate is based on the following: 1.

The ability to domain the resource rigorously by geology and therefore have a higher confidence in individual block estimates. Previous resource estimate had treated the resource, from an estimation viewpoint, as a single global domain.

2.

The use of a geologically-constrained LeapfrogTM wireframe grade shells to ensure the estimation was carried out within and at an appropriate cut-off grade and no dilution was allowed from low grade material outside this zone.

3.

The use of stratigraphic grids to ensure estimation of grade along the most geologically appropriate horizons.

4.

The availability of over 200 new density measurements, including bulk densities for the calcrete material, for most rock types within the resource. These data along with the use of a geologically-constrained model have improved the reliability of tonnage estimates. Previous estimates only used a single global density value.

5.

Continuity of mineralisation within the resource at the preferred cut-off grade of 100 ppm eU3O8. Sensitivity analyses of estimates by removing every second drillhole have shown little change in overall grade or tonnes. SRK therefore considers that the current average drill spacing of 100 x 100 m is sufficient to have a relatively high degree of confidence in block estimates of a similar size within the resource based on this low cut-off grade.

All lithologies which had no specific density information and where a global value of 1.76 t/m3 was used, were assigned to the Inferred category. These lithologies are: Mudstone Black shale Kopi Ferricrete Silcrete Carbonate rock Sand These lithologies accounted for less than 10% of the total resource by ore tonnage. All other blocks were assigned to the Indicated category. Drilling is of such a regular/even nature in the main resource area – approximately 100 x 100 m, even less in places – that there are no areas in which the estimate has taken place where the sample spacing is considerably more than this distance. Classification was based on the JORC Code for Reporting Mineral Resources (2004 edition) and is fully compliant with CIMM Guidelines for the Reporting of Resources.

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It is SRK‟s opinion that the improved geological domaining, significant new density data for the major rock types, sampling density and proved continuity of mineralisation allow up to 90% of the resource to be re-classified from Inferred to Indicated with the remaining 10% remaining in the Inferred category. It is recognised that infill drilling, additional density information and more mineralogical test work on recoveries will be required to convert parts of the resource to Measured status in the future. Once these issues have been reliably dealt with, it may be possible to upgrade parts of the resource to Measured.

19.4

Resource summary

The global NI43-101/JORC classified Resources are reported in Table 19-7 through Table 19-12 for various eU3O8 cut-offs globally and for the 100 ppm eU3O8 and 200 ppm eU3O8 cut-offs by lithology. Tonnage and grade figures are similar to those obtained in the 2007 estimate as few new holes have been drilled in the resource. However, the head grade at any particular cut-off is slightly higher due to more selective and geologically appropriate constraining of the estimate. Tonnages in some cases are lower due to the revised density values used which are based on the recent density measurements. Overall, the contained metal is slightly higher, up from 23 Mlbs in the earlier estimate to 26 Mlbs contained eU3O8. The most significant change in the current estimate is the upgrade of the majority of the resource from Inferred to Indicated category. At a 100 ppm eU3O8 lower cut-off grade, the vast majority of the uranium resource is hosted in the calcrete followed by the clay, sandy clay, and sand and calcrete clay material respectively. Whilst the Indicated Resource grade and tonnage are somewhat sensitive to increasing the cut-off grade from 100 ppm to 200 ppm eU3O8, this is matched by only a modest decrease in the reported contained metal. A decrease in the Indicated Resource tonnage of 9.85 Mt (34%) and increase in the Indicated Resource grade from 376 ppm eU3O8 to 497 ppm eU3O8 (32%) corresponds with a decrease in contained metal of 3.1 Mlb (13%) when raising the cut-off grade from 100 ppm eU3O8 to 200 ppm eU3O8. Table 19-7: Indicated Resource Cut-off grade (ppm eU3O8)

Ore Tonnage (kt)

Average grade (ppm eU3O8)

Contained eU3O8 (metal tonnes)

Contained eU3O8 6 (lbs metal x 10 )

100

28,751

376

10,810

23.83

150

23,445

426

9,987

22.01

200

18,901

497

9,394

20.71

250

14,976

569

8,521

18.78

500

6,077

882

5,360

11.81

Table 19-8: Inferred Resource Cut-off grade (ppm eU3O8)

Ore tonnage (kt)

Average grade (ppm eU3O8)

Contained eU3O8 (metal tonnes)

Contained eU3O8 (lbs metal x 106)

100

3,574

274

979

2.16

150

2,807

312

876

1.93

200

1,922

374

719

1.58

250

1,397

433

605

1.33

500

337

759

256

0.61

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Table 19-9: Indicated Resource at 100 ppm eU3O8 cut-off – by lithology Lithology

Ore tonnage (kt)

Grade (ppm eU3O8)

Metal tonnes (eU3O8)

Calcrete

11,039

511

5,641

Clay

8,856

325

2,878

Sandy Clay/Clayey Sand

6,834

254

1,736

Sandy

880

210

185

Calcrete clay

1,141

338

386

Total

28,751

376

10,810

Table 19-10: Indicated Resource at 200 ppm eU3O8 cut-off – by lithology Lithology

Ore tonnage (kt)

Grade (ppm eU3O8)

Metal tonnes (eU3O8)

Calcrete

8,929

591

5,277

Clay

5,538

449

2,487

Sandy Clay/Clayey Sand

3,157

372

1,174

Sandy

383

278

107

Calcrete clay

893

392

350

Total

18,901

497

9,394

Table 19-11: Inferred Resource at 100 ppm eU3O8 cut-off – by lithology Lithology

Ore tonnage (kt)

Grade (ppm eU3O8)

Metal tonnes (eU3O8)

Black shale

871

344

300

Mudstone

763

274

209

Carbonate rock

329

259

85

Gritty clay

130

314

41

Gravel

0

0

0

Kopi

0

0

0

Ferricrete

232

322

75

Evaporite Gypsum

1,177

215

253

Silcrete

73

246

18

Soil

0

0

0

Total

3,574

274

979

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Table 19-12: Inferred Resource at 200 ppm eU3O8 cut-off – by lithology Lithology

Ore tonnage (kt)

Grade (ppm eU3O8)

Metal tonnes (eU3O8)

Black shale

579

439

254

Mudstone

478

337

161

Carbonate rock

160

387

62

Gritty clay

113

334

38

Gravel

0

0

0

Kopi

0

0

0

Ferricrete

186

355

66

Evaporite Gypsum

367

348

128

Silcrete

39

298

12

Soil

0

0

0

Total

1,922

374

719

19.5

Model limitations

The current model has some important limitations at this point in time and these should be noted: Ongoing QA/QC on the gamma logs and routine comparisons with chemical assays are needed to ensure future sampling remains reliable More density information for the sand, carbonate rock, silcrete, ferricrete, kopi, mudstone and black shale units is required, so these can be reliably estimated in terms of tonnage Improvements in variography for all zones as more samples become available should increase the reliability of block estimates for future estimation upgrades Slope of regression analysis for the block models should be undertaken to determine reliability of estimates on a block by block basis for selective mining Selectivity – the model has a block size of 65mE x 85mN x 0.25 mRL and assumes mining selectivity to this level. Further mining studies are required to establish whether this is practical. It is highly likely that the mining method will be more selective than that, based on a grade control defined by radiometrics. If that is the case, there is a clear scope for a significant increase in the average grade, with some reduction in ore tonnage. To be able to estimate the potential gain of a more selective approach, kriging is not appropriate. Another non-linear technique, like Uniform Conditioning should be employed.

19.6

Recommendations

After the present re-estimation of the resources, SRK makes the following recommendations: Create a single database with different tables for repeats, duplicates, standards, chemical, gamma logs etc. More density data should be collected in future drilling programs to allow higher classifications to be assigned. Targeting density collection of under-sampled or unsampled rock types should be a priority.

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QA/QC on the gamma logs and chemical assays, as well as routine comparisons between radiometrics and chemical assays are needed to ensure future sampling remains reliable. To date no additional disequilibrium work has been undertaken on the latest Mega drilling. It is strongly recommended that some new disequilibrium studies be carried out as a routine check on at least 50 or so of the drill samples taken in the last three years from a variety of different geographic locations and rock types within the resource. Kriging should be accompanied by systematic tests of its quality (Quantitative Kriging neighbourhood Analysis) in order to be optimised. As it is likely that mining selectivity will be higher than the one implied by the current model, Mega should perform an estimation using a non-linear estimation method, like Uniform Conditioning, to evaluate the impact of better selectivity on ore tonnage and grade.

19.7

Modifying factors

Environmental, permitting, legal and title are discussed in Section 6 and the results of metallurgical testing to date are discussed in Section 18. To the best of the authors‟ knowledge there are no known environmental, permitting, legal, title, taxation, socio-economic, marketing, political, infrastructure or other issues that may materially affect the mineral resources at Lake Maitland.

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Other Relevant Data and Information

The authors of this technical report are not aware of additional information or explanations that, if excluded, would mislead the reader.

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Interpretation and Conclusions

Since the previous Technical Report in 2007 Mega have continued to evaluate the Lake Maitland uranium resource. An additional 794 aircore holes for a total of 11,677 m were drilled to bring drill spacing to a nominal 100 mN x 100mE spacing over the majority of the resource with the aim of improving the resource classification from Inferred to Indicated. Additional bulk density measurements were also made. The application of a 3D lithofacies model allowed application of these new data to provide greater confidence in the estimated tonnages. Perhaps the most significant change from previous estimates is the exclusion of all pre 2005 (pre-Redport) data from the estimation. It was recommended in the previous Technical Report that validation and QA/QC checks should be performed on the historical data. The additional drilling by Mega has in effect replaced these data, making these recommendations redundant. In conclusion, the latest Lake Maitland Resource estimate is considered to be sufficiently well sampled, geologically constrained, has demonstrated grade continuity for many of the geological host lithologies, and is sufficiently well tested in terms of density for accurate tonnage determinations to be made. In addition, no historic data have been used and the resource is based only on the more recent drilling campaigns of Redport in 2005 and Mega 2007/08 whose reliability has been examined and can be assured. This has resulted in the majority of the resource being reclassified from the Inferred to Indicated category. The current resource of the LMUP is 18.9 Mt at 497 ppm eU3O8 (Indicated) and 1.9 Mt at 374 ppm eU3O8 (Inferred) at a cut-off of 200 ppm eU3O8. The re-classification places approximately 90% of the resource into the Indicated category with 10% in the Inferred category. The 2007/2008 drilling programme by Mega has therefore met its objective to improve the classification. The highly geologically-constrained estimate resulted in a slight increase in average head grade for all cut-off ranges by approximately 13%. However, this was partly offset by the new density data resulting in some slightly lower densities than previously recorded for many of the main lithologies. Since the previous 2007 estimate, only minor drilling outside of the previous resource has taken place, and as a result the overall tonnage increase is small. By improving the classification of the mineral resource to Indicated, it now becomes possible to assess the project’s economic potential at a reasonable level of precision. This assessment is likely to take the form of a Feasibility Study.

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Recommendations

Recommendations pertaining to Metallurgy and Resource Estimation are given as part of sections 18 and 19 respectively. Given the significant upgrade in resource classification from Inferred to Indicated, the authors recommend that the LMUP proceed to Feasibility Study at the earliest opportunity. This recommendation is in line with Mega‟s stated objective of achieving production at Lake Maitland in Q4, 2011.

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References

Chiles, J P, Aug, C, Guillen, A and Lees, T, 2004. Modelling the Geometry of Geological Units and its uncertainty in 3D From Structural Data: The Potential-Field Method, paper presented to Orebody Modelling and Strategy Mine Planning Conference, Perth, Western Australia, 22-24 November. Conaway, J G, and Killeen, P G, 1978. Quantitative Uranium Determinations from Gamma-ray Logs by Application of Digital Times Series Analysis, Geophysics, 43, pp 1204-1221. Cowan, E J, Beatson, R K, Ross, H J, Fright, W R, McLennan, T J, Evans, T R, Carr, J C, Lane, R G, Bright, D V, Gillman, A J, Oshust, P A and Titley, M, 2003. Practical Implicit Geological Modelling, paper presented to 5th International Mining Geology Conference, Bendigo, Victoria, 17-19 November. Gleeson, P, 2009. Lake Maitland Resource Estimate, unpublished report prepared by SRK Consulting for Mega Uranium Ltd. Hellman and Schofield, 2007. First time disclosure: Mega Uranium Ltd. Mineral Resources for Lake Maitland Uranium Deposit. Technical Report (NI 43-101) Available through SEDAR website. McKay, A D and Miezitis, Y, 2001. Australia‟s uranium resources, geology and development of deposits, AGSO – Geoscience Australia, Mineral Resource Report 1. Princep D, 2006. Resources Estimation within the Lake Maitland Project, Western Australia, unpublished report prepared by Hellman and Schofield for Redport Ltd. Wilson D, 2006. The Application of Total Count Radiometric Logging to the Lake Maitland Uranium Deposit, Western Australia.

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Data and Signature Pages Certificate of Qualified Person Lake Maitland Western Australia I, Matthew Hayden Wheeler, do hereby certify that: 1.

I am the Manager Geology (Lake Maitland Project) of Mega Uranium Ltd of 57 Havelock Street, West Perth WA 6005, Australia. I have been an employee of Mega Uranium Ltd for a period of two and a half years.

2.

I am a graduate of Curtin University of Technology with a BSc in Applied Geology in 1994 and a BSc (Hons) in Applied Geology in 1996.

3.

I am a member of Australian Institute of Geoscientists, member number 3408.

4.

I have practiced my profession continuously since 1994.

5.

I have read the National Instrument 43-101 of the Canadian Securities Administrators („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 am a „qualified person‟ for the purposes of NI 43-101 and that the Technical Report has been prepared in compliance with this Instrument. I have contributed to Sections 5 to 15 and 18.2.1 of the technical report titled, “Lake Maitland – NI43-101 Technical Report” dated 8 September 2009 (the „Technical Report‟), relating to the Lake Maitland Uranium Project. I last visited the site on 20 March 2009 for a period of one (1) day.

6.

I have visited the Lake Maitland Uranium Project regularly since September 2006.

7.

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.

8.

I am an employee of Mega Uranium Ltd.

9.

I have read the National Instrument 43-101 and Form 43-101F1 (the „Form‟) and the Technical Report has been prepared in compliance with that Instrument and the Form.

Dated this 8 September 2009.

Matthew Wheeler BSc Hons (Applied Geology) MAIG Manager Geology (Lake Maitland Uranium Project)

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Certificate of Qualified Person Lake Maitland Western Australia

I, Daryl Peter Evans, do hereby certify that: 1.

I am the Principal Metallurgist of IMO Pty Ltd, a metallurgical consultancy. I have been an employee of IMO Pty Ltd for a period of two years, IMO are located at 88 Thomas Street West Perth, Western Australia.

2.

I am a graduate of Murdoch University with a Bachelor of Science degree (1986).

3.

I am a member of Australian Institute of Mining and Metallurgy, member number 224540.

4.

I have worked in the mining industry for over 23 years. Experience includes operational, technical and managerial roles in the: Gold, Iron Ore and Uranium sectors.

5.

I have read the National Instrument 43-101 of the Canadian Securities Administrators („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 am a „qualified person‟ for the purposes of NI 43-101 and that the Technical Report has been prepared in compliance with this Instrument. I am responsible for the information contained in Section 18 of the technical report titled, “Lake Maitland – NI43-101 Technical Report” dated 8 September 2009 (the „Technical Report‟), relating to the Lake Maitland Uranium Project. I last visited the site in March 2009 for a period of one (1) day.

6.

As Principal Metallurgist I have been responsible for the: scoping, design and development of metallurgical testwork the process flowsheet since January 2008.

7.

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.

8.

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

9.

I have read NI 43-101 and Form 43-101F1 and the Technical Report has been prepared in compliance with that instrument and form.

Dated this 8 September 2009.

Daryl Evans Principal Metallurgist IMO Pty Ltd

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Certificate of Qualified Person Lake Maitland Western Australia

I, Daniel Rene Guibal, FAusIMM (CP), do hereby certify that: 1.

I am a Corporate Consultant (Geostatistics & Resources) with SRK (Australasia) Pty Ltd, trading as SRK, which is an international firm of consulting geologists and engineers, which has been practicing in this profession since 1974. I hold office at 10 Richardson Street, West Perth, WA 6005, Australia and have been employed as such since 1999.

2.

I graduated from the Ecole Nationale Superieure de la Metallurgie et de l‟Industrie des Mines de Nancy, France with an Engineering Degree (Ingenieur Civil des Mines de Nancy, 1971) and I have continually practiced my profession since that time.

3.

I am a Fellow and Accredited Chartered Professional (Mining) of The Australasian Institute of Mining and Metallurgy, a Member of the Mineral Industry Consultants Association, Inc and a Life Member of the Geostatistical Association of Australasia.

4.

I have worked as a Mining Geostatistician for a total of 32 years since my graduation from University.

5.

I have read the definition of „qualified person‟ set out in National Instrument 43-101 of the Canadian Securities Administrators (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 am a „qualified person‟ for the purposes of NI 43-101. I am responsible for, with the exception of Section 18, the information contained in the technical report titled, “Lake Maitland – NI43-101 Technical Report” dated 8 September 2009 (the „Technical Report‟), relating to the Lake Maitland Uranium Project. I have not visited the site.

6.

My involvement in the project that is the subject of the Technical Report relates to the 2009 Resource Estimation and the preparation of this report.

7.

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.

8.

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

9.

I have read NI 43-101 and Form 43-101F1 and the Technical Report has been prepared in compliance with that instrument and form.

Dated this 8 September 2009.

Daniel R. Guibal Corporate Geostatistician, SRK Consulting

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Certificate of Qualified Person Lake Maitland Western Australia

I, Peter Gleeson, do hereby certify that: 1.

I am a Principal Consultant (Geology & Resources) with SRK (Australasia) Pty Ltd, trading as SRK, which is an international firm of consulting geologists and engineers, which has been practicing in this profession since 1974. I hold office at 10 Richardson Street, West Perth, WA 6005, Australia and have been employed as such since 2006.

2.

I graduated from the University of Leicester, UK with a Mining Geology Degree in 1981 and I have continually practiced my profession since that time. I completed an MSc in Applied Geostatistics/Mining Geology from Queensland University, Australia in 1992.

3.

I am a Member of the Australian Institute of Geoscientists

4.

I have worked as a Geologist for a total of 28 years since my graduation from University.

5.

I have read the definition of „qualified person‟ set out in National Instrument 43-101 of the Canadian Securities Administrators (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 am a „qualified person‟ for the purposes of NI 43-101. I am responsible for the information contained in Section 19 of the technical report titled, “Lake Maitland – NI43-101 Technical Report” dated 8 September 2009 (the „Technical Report‟), relating to the Lake Maitland Uranium Project. I visited the site for 2 days in March 2009.

6.

My involvement in the project that is the subject of the Technical Report relates to the 2009 Resource Estimation.

7.

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.

8.

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

9.

I have read NI 43-101 and Form 43-101F1 and the Technical Report has been prepared in compliance with that instrument and form.

Dated this 8 September 2009.

Peter Gleeson Principal Consultant, Geology, SRK Consulting

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Additional Requirements for Technical Reports on Development Properties and Production Properties

The Lake Maitland project is at an early stage of development as discussed in the technical report. There are no Additional Requirements for Technical Reports deemed applicable for this Technical Report in support of a revised resource estimate for the project.

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Appendices

Appendices

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Appendix 1

Appendix 1: Scatter plots of repeat downhole radiometric measurements

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Appendix 1-1

Lake Maitland - Repeat Gamma Logs

Lake Maitland - Repeat Gamma Logs

40.00

18.00 16.00

35.00 LMAC1044 y = 0.8767x + 2.5998 Repeat eU3O8 (ppm)

Repeat eU3O8 (ppm)

30.00

LMAC1045 y = 0.7964x + 2.6657

14.00

25.00 20.00 15.00 10.00

12.00

10.00 8.00 6.00 4.00

5.00

2.00

0.00

0.00 0.00

5.00

10.00

15.00

20.00

25.00

30.00

35.00

40.00

0.00

5.00

Original eU3O8 (ppm)

Lake Maitland - Repeat Gamma Logs

15.00

20.00

Lake Maitland - Repeat Gamma Logs

100.00

25.00

90.00

LMAC1046

80.00

LMAC1047 20.00

y = 1.0205x - 1.0366

70.00

Repeat eU3O8 (ppm)

Repeat eU3O8 (ppm)

10.00 Original eU3O8 (ppm)

60.00

50.00 40.00 30.00 20.00

y = 0.9144x + 1.0798

15.00

10.00

5.00

10.00 0.00

0.00 0.00

20.00

40.00

60.00

80.00

100.00

0.00

5.00

10.00

Original eU3O8 (ppm)

15.00

20.00

25.00

Original eU3O8 (ppm)

Lake Maitland - Repeat Gamma Logs

Lake Maitland - Repeat Gamma Logs

70.00

25.00 LMAC1048

LMAC1049

60.00

y = 0.7941x + 2.7384

20.00

y = 1.062x - 0.8269 Repeat eU3O8 (ppm)

Repeat eU3O8 (ppm)

50.00 40.00 30.00

15.00

10.00

20.00 5.00 10.00 0.00

0.00 0.00

10.00

20.00

30.00

40.00

50.00

60.00

70.00

0.00

5.00

10.00

Original eU3O8 (ppm)

15.00

20.00

25.00

Original eU3O8 (ppm)

Lake Maitland - Repeat Gamma Logs

Lake Maitland - Repeat Gamma Logs

25.00

35.00 LMAC1050

LMAC1051

30.00

y = 0.8016x + 1.77

20.00

y = 0.9642x + 1.7004 Repeat eU3O8 (ppm)

Repeat eU3O8 (ppm)

25.00 15.00

10.00

20.00 15.00 10.00

5.00 5.00 0.00

0.00 0.00

5.00

10.00

15.00

20.00

25.00

0.00

Original eU3O8 (ppm)

HERO/GLEE/GUIB/WILL/mool

5.00

10.00

15.00

20.00

25.00

30.00

35.00

Original eU3O8 (ppm)

MEG003_Lake_Maitland_NI43_101_Report_Rev9.doc

September 2009

SRK Consulting Project: MEG003 – Lake Maitland NI43-101 Report

Appendix 1-2

Lake Maitland - Repeat Gamma Logs

Lake Maitland - Repeat Gamma Logs

30.00

50.00 LMAC1052

25.00

20.00

Repeat eU3O8 (ppm)

Repeat eU3O8 (ppm)

y = 0.9049x + 0.921

15.00

10.00

45.00

LMAC1053

40.00

y = 0.9267x + 0.5305

35.00 30.00

25.00 20.00 15.00 10.00

5.00

5.00 0.00

0.00 0.00

5.00

10.00

15.00

20.00

25.00

30.00

0.00

10.00

20.00

Original eU3O8 (ppm)

Lake Maitland - Repeat Gamma Logs

40.00

50.00

60.00

Lake Maitland - Repeat Gamma Logs

18.00

800.00 LMAC1155

16.00

LMAC1162

700.00

y = 1.0014x - 0.0287

y = 0.6844x + 3.8085

14.00

600.00

12.00

Repeat eU3O8 (ppm)

Repeat eU3O8 (ppm)

30.00

Original eU3O8 (ppm)

10.00 8.00 6.00

500.00 400.00 300.00 200.00

4.00

100.00

2.00 0.00

0.00 0.00

2.00

4.00

6.00

8.00

10.00

12.00

14.00

16.00

0.00

100.00

200.00

Original eU3O8 (ppm)

300.00

400.00

500.00

600.00

700.00

800.00

Original eU3O8 (ppm)

Lake Maitland - Repeat Gamma Logs

Lake Maitland - Repeat Gamma Logs

600.00

70.00 LMAC1192

60.00

LMAC1163

500.00

y = 0.7711x + 4.4523 50.00 Repeat eU3O8 (ppm)

Repeat eU3O8 (ppm)

y = 0.9961x + 0.14 400.00

300.00

200.00

40.00 30.00 20.00

100.00

10.00

0.00

0.00 0.00

100.00

200.00

300.00

400.00

500.00

600.00

0.00

10.00

20.00

Original eU3O8 (ppm)

Lake Maitland - Repeat Gamma Logs

40.00

50.00

60.00

70.00

Lake Maitland - Repeat Gamma Logs

800.00

800.00

700.00

700.00

LMAC1222 y = 1.0069x - 2.893

500.00 400.00 300.00

500.00 400.00 300.00

200.00

200.00

100.00

100.00

0.00

LMAC1222 y = 1.0069x - 2.893

600.00 Repeat eU3O8 (ppm)

600.00 Repeat eU3O8 (ppm)

30.00

Original eU3O8 (ppm)

0.00 0.00

100.00

200.00

300.00

400.00

500.00

600.00

700.00

800.00

0.00

Original eU3O8 (ppm)

HERO/GLEE/GUIB/WILL/mool

100.00

200.00

300.00

400.00

500.00

600.00

700.00

800.00

Original eU3O8 (ppm)

MEG003_Lake_Maitland_NI43_101_Report_Rev9.doc

September 2009

SRK Consulting Project: MEG003 – Lake Maitland NI43-101 Report

Appendix 1-3

Lake Maitland - Repeat Gamma Logs

Lake Maitland - Repeat Gamma Logs

400.00

500.00

350.00

Repeat eU3O8 (ppm)

250.00 200.00 150.00

LMAC1226

400.00

y = 0.7296x + 17.932

300.00 Repeat eU3O8 (ppm)

450.00

LMAC1225

y = 1.0067x - 0.1262

350.00 300.00

250.00 200.00 150.00

100.00 100.00 50.00

50.00

0.00

0.00 0.00

50.00

100.00

150.00

200.00

250.00

300.00

350.00

400.00

0.00

100.00

Original eU3O8 (ppm)

300.00

400.00

500.00

Original eU3O8 (ppm)

Lake Maitland - Repeat Gamma Logs

Lake Maitland - Repeat Gamma Logs

500.00

300.00

450.00

LMAC1228

LMAC1227

400.00

250.00 y = 0.9969x + 0.6296

y = 1.0084x - 0.6788

350.00

Repeat eU3O8 (ppm)

Repeat eU3O8 (ppm)

200.00

300.00

250.00 200.00 150.00 100.00

200.00

150.00

100.00

50.00

50.00 0.00

0.00 0.00

100.00

200.00

300.00

400.00

500.00

0.00

50.00

Original eU3O8 (ppm)

100.00

150.00

200.00

250.00

300.00

Original eU3O8 (ppm)

Lake Maitland - Repeat Gamma Logs

Lake Maitland - Repeat Gamma Logs

350.00

600.00 LMAC1229

300.00

LMAC1230

500.00

y = 1.0197x - 0.1513

y = 0.9951x - 0.1055

Repeat eU3O8 (ppm)

Repeat eU3O8 (ppm)

250.00 200.00 150.00

400.00

300.00

200.00

100.00 100.00

50.00 0.00

0.00 0.00

50.00

100.00

150.00

200.00

250.00

300.00

350.00

0.00

Original eU3O8 (ppm)

100.00

200.00

300.00

400.00

500.00

600.00

Original eU3O8 (ppm)

Lake Maitland - Repeat Gamma Logs 600.00 LMAC1231

500.00

Repeat eU3O8 (ppm)

y = 1.021x - 1.2661 400.00

300.00

200.00

100.00

0.00 0.00

100.00

200.00

300.00

400.00

500.00

Original eU3O8 (ppm)

Scatter plot of radiometric repeat data by hole

HERO/GLEE/GUIB/WILL/mool

MEG003_Lake_Maitland_NI43_101_Report_Rev9.doc

September 2009

SRK Consulting Project: MEG003 – Lake Maitland NI43-101 Report

Appendix 2

Appendix 2: Twinned drillholes downhole graphs

HERO/GLEE/GUIB/WILL/mool

MEG003_Lake_Maitland_NI43_101_Report_Rev9.doc

September 2009

SRK Consulting Project: MEG003 – Lake Maitland NI43-101 Report

Appendix 2-1

Hole

X

Y

Z

Year

LMAC0080

307998.79

6991579.39

471.336

2005

LMAC1402

308012.54

6991577.12

471.38

2008

Lake Maitland twinned holes LMA0080/LMAC 1042 LMAC0080/LMAC 1042 120

100

80

LMAC0080

eU3O8

LMAC0080

dU3O8

LMAC1402

eU3O8

LMAC1402

dU3O8

eU3O8

60

40

20

0 0

5

10

15

20

25

Downhole depth

HERO/GLEE/GUIB/WILL/mool

MEG003_Lake_Maitland_NI43_101_Report_Rev9.doc

September 2009

SRK Consulting Project: MEG003 – Lake Maitland NI43-101 Report

Appendix 2-2

Hole

X

Y

Z

Year

LMAC0078

308397.1

6991578

471.55

2005

LMAC1056

308404.6

6991578

471.5

2008

LMSC024

308399.1

6991579

471.6

2008

Lake Maitland twinned holes LMAC0078/LMAC 1056/LMSC024 LMA0078/LMAC 1056/LMSC024 1200

1000

eU3O8

800

LMAC0078

U3O8

LMAC0078

eU3O8

LMAC0078

dU3O8

LMAC1056

eU3O8

LMAC1056

dU3O8

LMSC024

e 3O8

LMSC024

dU3O8

600

400

200

0 0

2

4

6

8

10

12

14

16

18

Downhole depth

HERO/GLEE/GUIB/WILL/mool

MEG003_Lake_Maitland_NI43_101_Report_Rev9.doc

September 2009

SRK Consulting Project: MEG003 – Lake Maitland NI43-101 Report

Appendix 2-3

Hole

X

Y

Z

Year

LMAC1039

309251.11

6990778.95

471.17

2008

LMAC1141

309248.36

6990773.82

471.12

2008

Lake Maitland twinned holes LMA1039/LMAC LMAC1039/LMAC1141 1411 350

300

250

eU3O8

200

LMAC1039

eU3O8

LMAC1039

dU3O8

LMAC1141

eU3O8

LMAC1141

dU3O8

150

100

50

0 0

2

4

6

8

10

12

14

Downhole depth

HERO/GLEE/GUIB/WILL/mool

MEG003_Lake_Maitland_NI43_101_Report_Rev9.doc

September 2009

SRK Consulting Project: MEG003 – Lake Maitland NI43-101 Report

Appendix 2-4

Hole

X

Y

Z

Year

LMAC0065

310003.17

6991579.54

471.43

2005

LMAC0961

309998.81

6991579.79

471.379

2008

Lake Maitland twinned holes LMAC0065/LMAC0961 3000

2500

eU3O8

2000

LMAC0065

U3O8

LMAC0065

eU3O8

LMAC0065

dU3O8

LMAC0961

eU3O8

LMAC0961

dU3O8

1500

1000

500

0 0

2

4

6

8

10

12

14

-500

Downhole depth

HERO/GLEE/GUIB/WILL/mool

MEG003_Lake_Maitland_NI43_101_Report_Rev9.doc

September 2009

SRK Consulting Project: MEG003 – Lake Maitland NI43-101 Report

Appendix 2-5

Hole

X

Y

Z

Year

LMAC0466

310474.8

6991784

470.854

2005

LMSC020

310469

6991784

470.84

2008

LMSC021

310477.4

6991784

470.89

2008

Lake Maitland twinned holes LMAC0466/LMSC020/LMSC021 600

500

eU3O8

400

LMAC0466

U3O8

LMAC0466

eU3O8

LMAC0466

dU3O8

LMSC020

eU3O8

LMSC020

dU3O8

LMSC021

eU3O8

LMSC021

dU3O8

300

eU3O8

200

100

0 0

5

10

15

20

25

30

Downhole depth

HERO/GLEE/GUIB/WILL/mool

MEG003_Lake_Maitland_NI43_101_Report_Rev9.doc

September 2009

SRK Consulting Project: MEG003 – Lake Maitland NI43-101 Report

Appendix 2-6

Hole

X

Y

Z

Year

LMAC0494

307801.8

6992281

472.745

2005

LMSC041

307797.1

6992281

472.8

2008

LMSC042

307803.8

6992282

472.79

2008

Lake Maitland twinned holes LMAC0494/LMSC041/LMSC042 350

300

250

eU3O8

200

LMAC0494

eU3O8

LMAC0494

dU3O8

LMSC041

eU3O8

LMSC041

dU3O8

LMSC042

eU3O8

LMSC042

dU3O8

150

100

50

0 0

1

2

3

4

5

6

7

8

9

10

Downhole depth

HERO/GLEE/GUIB/WILL/mool

MEG003_Lake_Maitland_NI43_101_Report_Rev9.doc

September 2009

SRK Consulting Project: MEG003 – Lake Maitland NI43-101 Report

Appendix 2-7

Hole

X

Y

Z

Year

LMAC0183

310949.18

6992380.53

470.921

2005

LMAC0893

310944.12

6992380.38

471

2008

Lake Maitland twinned holes LMAC04183/LMAC0893 1200

1000

eU3O8

800

LMAC0183

eU3O8

LMAC0183

dU3O8

LMAC0893

eU3O8

LMAC0893

dU3O8

600

400

200

0 0

2

4

6

8

10

12

Downhole depth

HERO/GLEE/GUIB/WILL/mool

MEG003_Lake_Maitland_NI43_101_Report_Rev9.doc

September 2009

SRK Consulting Project: MEG003 – Lake Maitland NI43-101 Report

Appendix 2-8

Hole

X

Y

Z

Year

LMAC0454

311392.76

6993372

470.176

2005

LMSC033

311388.16

6993373

470.3

2008

LMSC034

311396.15

6993372

470.17

2008

LMSC035

311393

6993367

470

2008

Lake Maitland twinned holes LMAC0454/LMSC033/LMSC034/LMSC035 3000

2500

eU3O8

2000

1500

LMAC0454

U3O8

LMAC0454

eU3O8

LMAC0454

dU3O8

LMSC033

eU3O8

LMSC033

dU3O8

LMSC034

eU3O8

LMSC034

dU3O8

LMSC035

eU3O8

LMSC035

dU3O8

1000

500

0 0

2

4

6

8

10

12

Downhole depth

HERO/GLEE/GUIB/WILL/mool

MEG003_Lake_Maitland_NI43_101_Report_Rev9.doc

September 2009

SRK Consulting Project: MEG003 – Lake Maitland NI43-101 Report

Appendix 2-9

Hole

X

Y

Z

Year

LMAC0198

310890.8

6992779

470.917

2005

LMSC018

310886.7

6992779

470.88

2008

LMSC019

310893

6992779

470.92

2008

Lake Maitland twinned holes LMAC0198/LMSC018/LMSC019 2500

2000

LMAC0198

U3O8

LMAC0198

eU3O8

LMAC0198

dU3O8

LMSC018

eU3O8

LMSC018

dU3O8

LMSC019

eU3O8

LMSC019

dU3O8

eU3O8

1500

1000

500

0 0

5

10

15

20

25

30

Downhole depth

HERO/GLEE/GUIB/WILL/mool

MEG003_Lake_Maitland_NI43_101_Report_Rev9.doc

September 2009

SRK Consulting Project: MEG003 – Lake Maitland NI43-101 Report

Appendix 2-10

Hole

X

Y

Z

Year

LMAC0722

311149.59

6994181.38

470.86

2008l

LMAC1397

311159.7

6994182.54

470.82

2008

Lake Maitland twinned holes LMAC0722/LMAC1397 1400

1200

1000

eU3O8

800

LMAC0722

eU3O8

LMAC0722

dU3O8

LMAC1397

eU3O8

LMAC1397

dU3O8

600

400

200

0 0

2

4

6

8

10

12

14

16

18

Downhole depth

HERO/GLEE/GUIB/WILL/mool

MEG003_Lake_Maitland_NI43_101_Report_Rev9.doc

September 2009

SRK Consulting Project: MEG003 – Lake Maitland NI43-101 Report

Appendix 2-11

Hole

X

Y

Z

Year

LMAC0505

309007.35

6993181.41

474.272

2005

LMSC039

309006.25

6993181.01

474.39

2008

Lake Maitland twinned holes LMAC0505/LMSC039 2500

2000

1500

LMAC0505

U3O8

LMAC0505

eU3O8

LMAC0505

dU3O8

LMSC039

eU3O8

LMSC039

dU3O8

eU3O8

1000

500

0 0

2

4

6

8

10

12

14

16

-500

Downhole depth

HERO/GLEE/GUIB/WILL/mool

MEG003_Lake_Maitland_NI43_101_Report_Rev9.doc

September 2009

SRK Consulting Project: MEG003 – Lake Maitland NI43-101 Report

Appendix 2-12

Hole

X

Y

Z

Year

LMAC0308

310600

6994782

470.732

2005

LMSC028

310594.8

6994782

470.67

2008

LMSC029

310601.9

6994781

470.65

2008

Lake Maitland twinned holes LMAC0308/LMSC028/LMSC029 4500

4000 LMAC0308 3500

3000

eU3O8

2500

U3O8

LMAC0308

eU3O8

LMAC0308

dU3O8

LMSC028

eU3O8

LMSC028

dU3O8

LMSC029

U3 O 8

LMSC029

dU3O8

2000

1500

1000

500

0 -5

0

5

10

15

20

25

-500

Downhole depth

HERO/GLEE/GUIB/WILL/mool

MEG003_Lake_Maitland_NI43_101_Report_Rev9.doc

September 2009

SRK Consulting Project: MEG003 – Lake Maitland NI43-101 Report

Appendix 2-13

Hole

X

Y

Z

Year

LMAC0386

311500.1

6995979

471.123

2005

LMSC001

311502.1

6995979

471.15

2008

LMSC002

311494.2

6995980

471.13

2008

Lake Maitland twinned holes LMAC0386/LMSC001/LMSC002 300

250

eU3O8

200

LMAC0386

U3O8

LMAC0386

dU3O8

LMSC001

eU3O8

LMSC001

dU3O8

LMSC002

eU3O8

LMSC002

dU3O8

150

100

50

0 0

2

4

6

8

10

12

14

Downhole depth

HERO/GLEE/GUIB/WILL/mool

MEG003_Lake_Maitland_NI43_101_Report_Rev9.doc

September 2009

SRK Consulting Project: MEG003 – Lake Maitland NI43-101 Report

Appendix 2-14

Hole

X

Y

Z

Year

LMAC0423

308500.4

6995977

471.388

2005

LMSC005

308495.54

6995977

471.44

2008

LMSC006

308502.61

6995978

471.45

2008

LMSC007

308502.41

6995972

471.6

2008

Lake Maitland twinned holes LMAC0423/LMSC005/LMSC006/LMSC007 400

350

300

eU3O8

250

200

LMAC0423

U3O8

LMAC0423

eU3O8

LMAC0423

dU3O8

LMSC005

eU3O8

LMSC005

dU3O8

LMSC006

U3 O 8

LMSC006

dU3O8

LMSC006

FROM

LMSC007

eU3O8

150

100

50

0 0

5

10

15

20

25

30

Downhole depth

HERO/GLEE/GUIB/WILL/mool

MEG003_Lake_Maitland_NI43_101_Report_Rev9.doc

September 2009

SRK Consulting Project: MEG003 – Lake Maitland NI43-101 Report

Appendix 2-15

Hole

X

Y

Z

Year

LMAC0403

311799.12

6996779.73

471.681

2005

LMAC1380

311786.9

6996777.32

471.74

2008

Lake Maitland twinned holes LMAC0403/LMAC1380 35

30

LMAC0403

eU3O8

LMAC0403

dU3O8

LMAC1380

eU3O8

LMAC1380

dU3O8

25

eU3O8

20

15

10

5

0 0

2

4

6

8

10

12

14

Downhole depth

HERO/GLEE/GUIB/WILL/mool

MEG003_Lake_Maitland_NI43_101_Report_Rev9.doc

September 2009

SRK Consulting Project: MEG003 – Lake Maitland NI43-101 Report

Appendix 2-16

Hole

X

Y

Z

Year

LMAC0444

307500.4

6996779

471.181

2005

LMSC003

307502.2

6996780

471.26

2008

LMSC004

307495.3

6996779

471.28

2008

Lake Maitland twinned holes LMAC0444/LMSC003/LMSC004 1000 900 800 700 600

LMAC0444

U3O8

LMAC0444

eU3O8

LMAC0403

dU3O8

LMSC003

eU3O8

LMSC003

dU3O8

LMSC004

eU3O8

LMSC004

dU3O8

eU3O8

500 400 300 200 100 0 0

2

4

6

8

10

12

14

-100

Downhole depth

HERO/GLEE/GUIB/WILL/mool

MEG003_Lake_Maitland_NI43_101_Report_Rev9.doc

September 2009

SRK Consulting Project: MEG003 – Lake Maitland NI43-101 Report

Appendix 3

Appendix 3: Apparent SG determinations on dry calcrete rock specimens

HERO/GLEE/GUIB/WILL/mool

MEG003_Lake_Maitland_NI43_101_Report_Rev9.doc

September 2009

SRK Consulting Project: MEG003 – Lake Maitland NI43-101 Report

Area

CEC-2

CEC-4

CEC-5

CEC-6

CEC-7

CEC-8

CEC-9

CEC-10

CEC-11

CEC-12

HERO/GLEE/GUIB/WILL/mool

Appendix 3-1

Sample number

Dry weight (g)

Coated weight (g)

Immersed weight (g)

Apparent SG

1

542.1

550.5

293.9

2.20

2

457.8

466.2

247.6

2.19

3

574.0

587.4

292.9

2.06

4

740.7

760.6

400.8

2.20

5

554.6

566.4

260.8

1.90

6

523.5

532.1

265.4

2.04

1

265.4

275.8

149.3

2.32

2

912.9

929.2

460.2

2.03

3

1300.1

1321.3

739.6

2.34

4

1335.2

1358.4

725.4

2.20

5

1080.5

1107.4

630.9

2.43

6

483.7

498.2

270.1

2.29

1

285.7

295.5

158.9

2.28

2

1121.3

1144.3

563.6

2.03

3

420.5

431.1

241.6

2.38

4

515.9

528.4

258.1

2.02

5

1121.6

1149.8

634.2

2.32

6

253.4

261.2

126.7

2.02

1

1072.8

1095.3

516.4

1.94

2

94.4

97.5

45.7

1.96

3

593.7

607.6

283.6

1.93

4

453.2

463.7

230.3

2.05

5

385.6

396.1

170.0

1.80

6

395.1

406.6

186.4

1.91

1

1988.7

2016.6

1178.5

2.47

1

328.1

337.0

188.4

2.38

2

2432.2

2464.5

1298.7

2.16

3

2209.4

2249.6

1230.1

2.27

4

1714.3

1743.6

966.5

2.31

5

1150.3

1170.7

671.6

2.42

6

1436.5

1458.2

809.4

2.30

1

565.5

585.0

306.2

2.21

2

449.5

462.3

228.7

2.06

1

315.8

323.3

142.7

1.84

2

367.0

377.4

192.8

2.13

3

769.7

788.5

414.7

2.19

4

246.4

253.5

137.0

2.28

5

1666.8

1701.9

914.8

2.23

6

580.1

599.2

276.2

1.93

1

1101.6

1127.4

591.8

2.18

2

1747.2

1776.7

910.2

2.10

3

490.8

511.6

241.9

2.00

4

297.5

305.1

142.6

1.94

1

1136.1

1158.3

628.6

2.26

MEG003_Lake_Maitland_NI43_101_Report_Rev9.doc

September 2009

SRK Consulting Project: MEG003 – Lake Maitland NI43-101 Report

Area

CEC-13

CEC-14

CEC-15

HERO/GLEE/GUIB/WILL/mool

Appendix 3-2

Sample number

Dry weight (g)

Coated weight (g)

Immersed weight (g)

Apparent SG

2

463.2

473.7

270.0

2.42

3

150.9

156.4

65.4

1.79

4

729.7

750.2

403.8

2.26

5

1047.2

1069.3

607.4

2.40

6

751.8

765.6

411.3

2.22

1

723.9

738.4

377.1

2.10

2

3060.1

3101.5

1750.5

2.35

3

211.8

216.7

110.7

2.11

4

878.9

897.6

443.4

2.03

5

543.5

554.3

252.1

1.88

6

2066.0

2098.7

1044.2

2.03

1

960.8

1000.3

527.1

2.25

2

950.3

968.0

517.5

2.21

3

1138.9

1156.5

644.3

2.32

4

1061.1

1076.6

611.2

2.37

5

1246.7

1271.6

685.3

2.24

6

484.9

499.5

280.0

2.40

1

2514.5

2574.7

1450.2

2.39

2

1643.7

1677.0

965.4

2.44

3

1175.3

1200.9

683.6

2.41

4

483.0

492.7

281.1

2.41

5

760.5

784.1

446.5

2.45

6

3518.3

3591.1

2094.9

2.49

MEG003_Lake_Maitland_NI43_101_Report_Rev9.doc

September 2009