national instrument ni 43-101 report u3o8 corp. · 1 berlin project, colombia national instrument...

1 BERLIN PROJECT, COLOMBIA

National Instrument NI 43-101 Report

Prepared by Coffey Mining Pty Ltd on behalf of:

U3O8 Corp.

Effective Date: 2 March 2012

Qualified Person: Neil Inwood - BSc (Geol.), PGradDip (Geol), MSc, FAusIMM John Goode - BSc., P.Eng.,FAusIMM, MCIM Paul Miller - Ph.D,MIMM, Chart. Eng

MINEWPER00790AC

Coffey Mining Pty Ltd

DOCUMENT INFORMATION

Berlin Project, Colombia – MINEWPER00790AC National Instrument NI 43-101 Report – 2 March 2012

Author(s): Neil Inwood Principle Resource Geologist BSc (Geol.), PGradDip (Geol), MSc, FAusIMM

John Goode Consulting Metallurgist BSc., P.Eng.,FAusIMM, MCIM

Paul Miller Manging Director Ph.D,MIMM, Chart. Eng.

Date: 2 March 2012

Project Number: MINEWPER00790AC

Version / Status: Final

Path & File Name: F:\MINE\Projects\U3O8 Corp.\MINEWPER00790AC_Berin Project 2011\Report Preparation\Text\CMWPr_790AC_Berlin_43-101_Feb2012_Final.docx

Print Date: Friday, 2 March 2012

Copies: U3O8 Corp.oration (2)

Coffey Mining – Perth (1)

Document Change Control

Version Description (section(s) amended) Author(s) Date

Document Review and Sign Off

[signed] [signed]

Primary Author Neil Inwood

Supervising Principal Ingvar Kirchner



DATE AND SIGNATURE PAGE

The “qualified persons” (within the meaning of NI 43-101) for the purposes of this report are

Mr Neil Inwood, Mr John Goode and Dr Paul Miller. The effective date of this report is March 2, 2012.

[signed] Neil Inwood BSc (Geol.), PGradDip (Geol), MSc, FAusIMM

Principle Resource Geologist Coffey Mining Pty Ltd. Signed on the 2 March 2012

[signed] Mr John Goode B.Sc., P.Eng., FAusIMM, MCIMM

Consulting Metallurgist J.R. Goode and Associates Signed on the 2 March 2012

[signed] Dr Paul Miller Ph.D, MIMM, Chart. Eng.

Managing Director Sulphide Resource Processing Pty Ltd Signed on the 2 March 2012



Table of Contents

Date and Signature Page ............................................................................................................................... i

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

2 Introduction ......................................................................................................................................... 3

2.1 Scope of Work .................................................................................................................................... 3

2.2 Principal Sources of Information ........................................................................................................ 3

2.3 Participants ......................................................................................................................................... 4

2.4 Site Visit .............................................................................................................................................. 4

2.5 Qualifications and Experience ............................................................................................................ 4

2.6 Independence ..................................................................................................................................... 5

2.7 Abbreviations ...................................................................................................................................... 5

3 Reliance on Other Experts ................................................................................................................. 7

4 Property Description and Location ................................................................................................... 8

4.1 Regulatory Framework for Mineral Properties in Colombia ............................................................... 8

4.1.1 Overview ........................................................................................................................... 8 4.1.2 Legal and Regulatory Framework of Mineral Concessions ............................................. 8 4.1.3 Repatriation of Funds and Payment of Dividends ............................................................ 9 4.1.4 Canada-Colombia Free Trade Agreement ....................................................................... 9 4.1.5 Mineral Concessions ......................................................................................................... 9 4.1.6 Other Required Permits and Environmental Liabilities .................................................. 10 4.1.7 Concession Fees ........................................................................................................... 11

4.2 Royalties .......................................................................................................................................... 11

4.3 Area of the Berlin Project Properties ............................................................................................... 11

4.4 Location of the concessions ............................................................................................................ 12

4.5 Location of Mineralised Zones Relative to the Properties .............................................................. 13

4.6 Details Pertaining to the Concessions ............................................................................................ 13

4.6.1 Concession 1 (File No 755-17) ...................................................................................... 13 4.6.2 Concession 2 (File No 756-17) ...................................................................................... 14 4.6.3 Concession 3 (File No 664-17) ...................................................................................... 15 4.6.4 Concession 4 (File No IFM 08221X) ............................................................................. 16 4.6.5 Concession 5 (File No 736-17) ...................................................................................... 17

4.7 Related Agreements ........................................................................................................................ 18

4.8 Comments ....................................................................................................................................... 18

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

5.1 Topography, Elevation and Vegetation ........................................................................................... 19

5.2 Access and Infrastructure ............................................................................................................... 20

5.3 Climate ............................................................................................................................................. 21

6 History ................................................................................................................................................ 22

6.1 Prior Ownership of the Berlin Property ........................................................................................... 22

6.2 Prior Exploration .............................................................................................................................. 22

6.3 Historic Resource Estimates ........................................................................................................... 28

6.3.1 Historic Resource Estimate ........................................................................................... 28 6.4 Metallurgical Testwork ..................................................................................................................... 28

7 Geological Setting and Mineralisation ............................................................................................ 29

7.1 Regional Geology ............................................................................................................................ 29



7.1.1 Lithology and Stratigraphy ............................................................................................. 29 7.1.2 Structure ......................................................................................................................... 33

7.2 Geology of the Berlin Project .......................................................................................................... 34

7.2.1 Lithology & Stratigraphy ................................................................................................. 34 7.2.2 Structure ......................................................................................................................... 40

7.3 Mineralisation .................................................................................................................................. 43

7.3.1 Method of Study ............................................................................................................. 43 7.3.2 Description of the Mineralised Zone .............................................................................. 43 7.3.3 Composition and Textures of the Host Rocks ............................................................... 44 7.3.4 Nature of the Mineralization ........................................................................................... 50 7.3.5 Alaskite ........................................................................................................................... 56 7.3.6 Summary of Mineral Textures & Observations Regarding Paragenesis ...................... 57

8 Deposit Types ................................................................................................................................... 59

8.1 Historic Perspective ......................................................................................................................... 59

8.2 Analogous Deposits ........................................................................................................................ 59

9 Exploration ........................................................................................................................................ 60

9.1 Responsibility for Exploration .......................................................................................................... 60

9.2 Approach ......................................................................................................................................... 60

9.3 Trenching ......................................................................................................................................... 60

9.4 Conclusion ....................................................................................................................................... 60

10 Drilling ................................................................................................................................................ 64

10.1 Drill Programs .................................................................................................................................. 64

10.2 Discussion of Drill Results ............................................................................................................... 69

11 Sample Preparation, Analyses and Security .................................................................................. 70

11.1 Qualification of Personnel and Responsibility for Sampling ........................................................... 70

11.2 Laboratory Certification ................................................................................................................... 70

11.3 Sampling Procedure ........................................................................................................................ 70

11.3.1 Exploration and Trench Samples ................................................................................... 70 11.3.2 Drill Core Samples ......................................................................................................... 71

11.4 Sample Preparation ......................................................................................................................... 74

11.5 Sample Analysis .............................................................................................................................. 74

11.5.1 ME-MS61U ..................................................................................................................... 74 11.5.2 ME-M61 .......................................................................................................................... 75 11.5.3 ME-MS81 ....................................................................................................................... 75 11.5.4 ME-XRF .......................................................................................................................... 75 11.5.5 AA24 ............................................................................................................................... 75

11.6 Comments ....................................................................................................................................... 75

12 Data Verification ................................................................................................................................ 76

12.1 Verification of Data .......................................................................................................................... 76

12.2 Independent Sampling .................................................................................................................... 76

12.3 Analytical Quality Control Procedure and Data .............................................................................. 76

12.3.1 U3O8 Corp. Submitted Standards ................................................................................. 78 12.3.2 Drillhole Pulp Duplicates ................................................................................................ 87

12.4 Comments ....................................................................................................................................... 89

13 Mineral Processing and Metallurgical Testing ............................................................................... 90

13.1 Approach ......................................................................................................................................... 90

13.2 Laboratories and Consultants ......................................................................................................... 90



13.3 Nature of Material ............................................................................................................................ 91

13.4 Head Grade & Composition of Composite Samples ...................................................................... 93

13.5 Metallurgical Testwork ..................................................................................................................... 95

13.5.1 Baseline Leach Tests on Raw (Unbeneficiated) Mineralised Material .......................... 95 13.5.2 Acetic Acid Pre-Leach Followed by Strong Acid Leach ................................................ 95 13.5.3 Ferric Leach ................................................................................................................... 98 13.5.4 Flotation Tests .............................................................................................................. 105

13.6 Summary and Conclusions ........................................................................................................... 110

14 Mineral Resource Estimates .......................................................................................................... 113

14.1 Resource Database and Validation .............................................................................................. 113

14.1.1 Database ...................................................................................................................... 113 14.1.2 Validation ...................................................................................................................... 113

14.2 Geological Interpretation and Modelling ....................................................................................... 114

14.2.1 Geological and Mineralisation Model ........................................................................... 114 14.3 Weathering and Topographic Profile ............................................................................................ 114

14.4 Statistical Analysis ......................................................................................................................... 117

14.4.1 Radiometric Data ......................................................................................................... 117 14.4.2 Summary Statistics and Top Cuts ............................................................................... 117 14.4.3 Bulk Density Data......................................................................................................... 118 14.4.4 Variography .................................................................................................................. 123

14.5 Block Model Construction ............................................................................................................. 126

14.6 Grade Estimation Parameters ....................................................................................................... 127

14.6.1 U3O8 Grade Estimate ................................................................................................... 127 14.6.2 Multi-Element Data ...................................................................................................... 128

14.7 Bulk Density ................................................................................................................................... 128

14.8 Resource Reporting and Classification ......................................................................................... 128

14.8.1 Introduction .................................................................................................................. 128 14.8.2 Criteria for Resource Categorisation ........................................................................... 129 14.8.3 Grade Tonnage Reporting ........................................................................................... 131

14.9 Conclusions ................................................................................................................................... 131

15 Mineral Reserve Estimates ............................................................................................................ 132

16 Mining Methods ............................................................................................................................... 133

17 Recovery Methods .......................................................................................................................... 134

18 Project Infrastructure ..................................................................................................................... 135

19 Market Studies and Contracts ....................................................................................................... 136

20 Environmental Studies, Permitting and Social or Community Impact ...................................... 137

21 Capital and Operating Costs .......................................................................................................... 138

22 Economic Analysis ......................................................................................................................... 139

23 Adjacent Properties ........................................................................................................................ 140

24 Other Relevant Data and Information ........................................................................................... 141

25 Interpretation and Conclusions ..................................................................................................... 142

26 Recommendations .......................................................................................................................... 143

27 References ....................................................................................................................................... 144



List of Tables

Table 1_1 – Berlin Project, Colombia – January 17 Resource Estimate, 2012 1

Table 2.7_1 – List of Abbreviations 6

Table 4.1.7_1 – Table of Annual Concession Fees 11

Table 4.2_1 – List of NSR on the Mining of Various Commodities 11

Table 4.4_1 – Corner Points of the Berlin Project Concessions 12

Table 6.2_1 –Rock-Chip Sample Assay Results from Radioactive Unit, Southern Berlin Project Area 25

Table 6.2_2 – Rock-Chip Sample U3O8 Assay Values and Thickness 25

Table 6.2_3 – Location of the Adit Portals Excavated by Minatome Exploration Program, Southern Berlin Project 27

Table 3.3_1 – Relative Concentrations of Minerals in the ANSTO1 Composite Sample of Carbonate Facies 47

Table 3.3_2 – Comparison of the Composition of the Two Types of Apatite in Carbonate Facies Rocks in Drillhole DDB7 49

Table 9.3_1 – Summary of Trench Assay Results 61

Table 10.1_1 – Summary of Drillhole Length and Drill Platform Number for Collars 66

Table 10.1_2 – Summary of Assay Results from 2010-2011 Drilling Program 68

Table 12.2_1 – Independent Sampling Results 77

Table 12.3.1_1 – U308 Corp Submitted Standards Expected Value for Main Elements 78

Table 12.3.1_2 –U308 Corp Submitted U Standards 79

Table 12.3.1_3 – Statistics for U308 Corp submitted Standards (V ppm) 82

Table 12.3.1_4 – U308 Corp Submitted Mo Standards 82

Table 12.3.1_5 –U308 Corp Submitted P Standards 84

Table 12.3.1_6 –U308 Corp Submitted Y Standards 84

Table 12.3.2_1 – U308 Corp Drillhole Pulp Duplicate Analysis (ALS Lab) 87

Table 12.3.2_2 – U308 Corp Trench Field Duplicate Analysis (ALS Lab) 88

Table 12.3.2_3 – Summary of Data Precision U308Corp Lab Duplicate Analysis (ALS Lab) 88

Table 13.3_1 – Details of Drillhole Intercepts Used in Metallurgical Testwork to Date 92

Table 13.4_1 – Head Grade of Composite Samples Used in Metallurgical Testwork by the Various Laboratories 94

Table 13.5.1_1 – Composite Sample BER-16061 - Summary of Alkaline Leach Conditions and Metal Extractions Achieved 96

Table 13.5.1_2 – Various Composite Samples - Summary of Alkaline Leach Conditions and Metal Extractions Achieved 96

Table 13.5.2_1 – Composite Sample BER-16061 - Summary of Conditions of Alkaline Leach and Metal Extractions 97

Table 13.5.1_1 – Summary of Ferric Leach Tests on Two Composite Samples From Berlin 99

Table 13.5.1_2 – Summary of Ferric Leach Tests Conditions (Without Stage 2 Leach) showing Effect of Temperature on Metal Recoveries 101

Table 13.5.3_3 – Uranium Extraction Obtained in Corroborative Ferric Leach Tests 101

Table 13.5.3_4 – Calculation of Reagent Consumption in the Two-Step Ferric Leach followed by Acid Wash Tests Undertaken 104

Table 13.5.4_1 – Summary of Carbon and Sulphide Pre-float and Apatite or Carbonate Flotation Tests Results 107

Table 13.5.4_2 – Summary of Sulphuric Acid Leach Test Results Undertaken at a Temperature of 65°C on Flotation Products 108

Table 14.4.2_1 – Summary Mineralised Zone Composite Statistics 119

Table 14.4.3_1 – Bulk Density Data Summary Statistics 123

Table 14.4.4_1 – Omnidirectional Variogram for 0.8m U3O8 (ppm) Composites 126

Table 14.5_1 – Block Model Parameters 127

Table 14.5_2 – Block Model Variables 127

Table 14.6.1_1 – U3O8 Sample Search Parameters – Ordinary Kriging 128

Table 14.6.2_1 – Multi-Element Sample Search Parameters – ID2 128

Table 14.8.2_1 – Confidence Levels of Key Categorisation Criteria 129

Table 14.8.3_1 – January 17 Resource Estimate, 2012,. 131

Table 26_1 – Costing for the Berlin Project Recommendations 143



List of Figures

Figure 4.5_1 – Location of the Berlin Project Exploration Concessions 13

Figure 5.2_1 – General Location of the Berlin Concession Areas in Caldas Province, Columbia 19

Figure 5.1_2 – Photo showing the Typical Landscape of the Berlin Project 20

Figure 5.2_2 – Google Satellite Image showing Secondary Road Access 21

Figure 6.2_1 – Channel Sample Uranium Grades in Uraniferous Black Shale, Northern Berlin Project Area 23

Figure 6.2_2 – Channel Sample Uranium Grades in Uraniferous Black Shale, Southern Berlin Project Area 24

Figure 6.2_3 – Location of the Adit Portals Excavated by Minatome Exploration Program, Southern Berlin Project 26

Figure 6.2_4 – Historical drilling in the Berlin Area Area 27

Figure 7.1_1 – Main Tectonic Components of Colombia 30

Figure 7.1_2 – Main Tectonic Components of Colombia 31

Figure 7.1_3 – Regional Geological Setting of the Berlin Project 31

Figure 7.2.1_1 – Geology of the Berlin Project 35

Figure 7.2.1_2 – Generalised Stratigraphic Column of the Berlin Project 36

Figure 7.2.1_3 – Generalised Stratigraphic Column of the Berlin Project 37

Figure 7.2.1_4 – Drillhole DDB-013 Sedimentary Sequence 38

Figure 7.2.2_1 – West-East Cross Sections (1-4) Through Berlin Syncline and Plan View 41

Figure 7.2.2_2 – West-East Cross Sections (5-12)Through Berlin Syncline 42

Figure 7.2.3_1 – Geological Profile of Trench TB28 Located on the West Limb of the Berlin Syncline on Section P1 43

Figure 7.3.3_1 – Sandstone Photomicrographs 45

Figure 7.3.3_2 – Sandstone Photomicrographs In Plane and Crossed Polarised Light 46

Figure 7.3.3_3 – Carbonate Facies Photomicrographs from a Clastic Layer 48

Figure 7.3.3_4 – Backscatter Images of Fractures 49

Figure 7.3.4_1 – Backscatter Images of Uranium Bearing Minerals 50

Figure 7.3.4_2 – BSE Image of the Distribution of Uraninite Particles at Grain Boundaries within the Carbonate Facies 51

Figure 7.3.4_3 – BSE Image Showing Pale Uranium Mineral Inclusions In Stage 2 Fluorapatite 51

Figure 7.3.4_4 – Graphite In Reflected Light Microscopy Images 52

Figure 7.3.4_5 – Evidence of Stage 2 Phosphate Precipitation in Transmitted Light Microscopy Image 53

Figure 7.3.4_6 – Backscatter Image of Branching Stringers of Calcite 53

Figure 7.3.4_7 – Backscatter Image of Replacement of Graphite by Calcite 54

Figure 7.3.4_8 – Backscatter Image of Replacement of Graphite by Calcite 55

Figure 7.3.4_9 – BSE Ni-As Sulphide (Gesdorffite?) in Contact with Quartz and Calcite from the Carbonate Facies 55

Figure 7.3.5_1 – Photomicrographs of Coarse Plagioclase 56

Figure 9.3_1 – Geological Map of Southern Berlin Syncline Showing Location of Trenches Completed in 2010 and 2011 62

Figure 9.3_2 – Geological Map of Northern Berlin Syncline Showing Location of Trenches Completed in 2010 & 2011 63

Figure 10.1_1 – Monthly Drilling Rate In Berlin 64

Figure 10.1_2 – Location of Drill Platforms 65

Figure 11.3.2_1 – Core Sampling and Storage 73

Figure 12.3.1_1 – Berlin Project Uranium Standard Control Plot 80

Figure 12.3.1_2 – Berlin Project Vanadium Standard Control Plots 81

Figure 12.3.1_3 – Berlin Project Molybdenum Standard Control Plots 83

Figure 12.3.1_4 – Berlin Project Phosphorus Standard Control Plots 85

Figure 12.3.1_5 – Berlin Project Phosphorus Standard Control Plot 86

Figure 12.3.1_6 – Berlin Project Yttrium Standard Control Plots 86

Figure 13.3_1 – Figure Description 93

Figure 13.5.4_1 – Leach Kinetics on Flotation Products from Composite Sample DDB10-15 109



Figure 13.5.4_2 – Plot Of Total Organic Carbon Recovery in Pre-Float Concentrate versus Uranium Recovery 109

Figure 14.2.1_1 – Modelled Mineralisation Used for 2012 Resource 115

Figure 14.2.1_2 – Oblique Section 618000mN with Drilling and U3O8 Values 116

Figure 14.4.1_1 – Scatter Plot of Chemical and Radiometric U3O8 Data 117

Figure 14.4.2_1 – Multi Element Histogram Plots – Mineralised Zone 120

Figure 14.4.2_2 – Scatter Plots for U3O8 Compared to Other Elements 121

Figure 14.4.2_3 – Scatter Plots for U3O8 Compared to Other Elements 122

Figure 14.4.3_1 – Berlin Density Histograms 124

Figure 14.4.4_1 – Omnidirectional U3O8 Variography 126

Figure 14.8.2_1 – Drillhole Locations, Mineralisation Model and Classification 130

List of Appendices

Appendix A – QAQC Summary Plots

Appendix B – Qualified Persons Certificates


Berlin Project, Colombia – MINEWPER00790AC Page: 1 National Instrument NI 43-101 Report – 2 March 2012

2 0B0BSUMMARY

This technical report summarises the resource estimation studies undertaken in January 2012

on U3O8 Corp.’s Berlin Property in Colombia. Mineral properties that constitute the Berlin

Project are owned by U3O8 Corp. through its wholly-owned subsidiary Gaia Energy

Investments Ltd. (formerly Energentia Ltd.). One of the properties (736-17) is in the process of

being transferred to Gaia Energy from Anglo Gold Ashanti Limited.

This report complies with disclosure and reporting requirements set forth in the TSX Venture

Exchange (TSXV) Corporate Finance Manual, National Instrument 43-101, Companion Policy

43-101CP, and Form 43-101F1.

The Berlin Project area is located in central Colombia in the province of Caldas some 80km

northeast of the provincial capital, Manizales and approximately 150km northeast of the national

capital, Bogotá. The Berlin Project lies within an area of 10,681 Hectares and is covered by

five contiguous concessions.

U3O8 Corp. is targeting uranium, vanadium, phosphorus, yttrium, molybdenum, nickel,

rhenium and neodymium mineralisation within a keel-shaped fold. The project is in steep

terrain in which drilling is conducted from platforms cut into hillsides and trenching is excavated

by hand in areas where the mineralisation comes to surface.

The maiden NI 43-101 compliant Resources estimate completed by Coffey Mining in January

2012 includes 8.1Mt at a grade of 0.11% U3O8 of Inferred Resources and 0.6Mt at a grade of

0.11% U3O8 Indicated Resources above a 0.04% U3O8 lower cutoff (Table 1_1). This updates

a non-NI 43-101 compliant, historical estimate undertaken by Minatome in 1981.

Table 1_1

Berlin Project, Colombia

January 17 Resource Estimate, 2012

Reported above various U3O8 lower cutoffs using a bulk density of 2.72t/m³ for fresh material and 2.0t/m³ for weathered material

Ordinary Kriged Estimate for U3O8, multi-element data estimated using Inverse Distance to the power of 2 using 0.8m assay composite data.

Parent Block of 50m(Y) x 4m (X) by 40m (Z) Preferred Reporting Cutoff – 0.04% U3O8

Lower Cutoff

(% U3O8) Mt

U3O8 %

Contained P2O5 (%)

V2O5

(%) Y2O3

(ppm) Mo

(ppm) Ni (%)

Ag (ppm)

Re (ppm)

Nd (ppm) U3O8

(Mkg) U3O8 (MLb)

Inferred

0.04 8.1 0.11 9.0 19.9 9.4 0.5 500 620 0.2 3.4 6.8 100

0.05 8.0 0.11 9.0 19.7 9.4 0.5 500 620 0.2 3.3 6.8 100

0.06 8.0 0.11 8.9 19.7 9.4 0.5 500 620 0.2 3.3 6.8 100

0.07 7.9 0.11 8.9 19.5 9.5 0.5 510 620 0.2 3.3 6.8 100

0.08 7.7 0.11 8.7 19.2 9.5 0.5 510 630 0.2 3.3 6.9 100

0.09 6.8 0.12 7.9 17.5 9.7 0.5 520 630 0.2 3.4 7.0 110

0.1 5.6 0.12 6.8 15.0 10.0 0.5 540 650 0.2 3.5 7.2 110

Indicated

0.04 0.6 0.11 0.7 1.5 8.4 0.4 460 570 0.2 2.8 6.1 110

0.05 0.6 0.11 0.7 1.5 8.4 0.4 460 570 0.2 2.8 6.1 110

0.06 0.6 0.11 0.7 1.5 8.4 0.4 460 570 0.2 2.8 6.1 110

0.07 0.6 0.11 0.7 1.5 8.4 0.4 460 570 0.2 2.8 6.1 110

0.08 0.6 0.11 0.7 1.5 8.4 0.4 460 570 0.2 2.8 6.1 110

0.09 0.6 0.11 0.7 1.5 8.4 0.4 460 580 0.2 2.9 6.1 110

0.1 0.5 0.11 0.6 1.2 8.6 0.4 480 590 0.2 2.9 6.3 110



The Berlin Project represents an exploration project with NI 43-101 compliant Indicated and

Inferred resources being defined for the first time.

Coffey Mining has reviewed the trenching, drilling, sampling, assaying and field procedures

used by U3O8 Corp. and consider them to be of overall high quality. Further infill drilling is

required to better define the mineralisation and define the shape of the fold structure in more

detail. Assaying of the drillholes should take into account both radiometric and chemical

assaying methods to allow for uranium grade estimation in intervals with poor drill core recovery.

Metallurgical testing has shown that uranium, phosphate, vanadium and the suite of other

metals of potential economic interest at Berlin is amenable to extraction via several pathways.

Of principal importance is ferric iron leach followed by a dilute acid wash from which excellent

rates of extraction were obtained from unbeneficiated ore. Further testwork is underway to

refine this extraction process. A second potential extraction route involves beneficiation of the

ore by flotation of organic carbon (mainly graphite) with the majority of the uranium

mineralisation, followed by acid leach to extract the uranium. Flotation tests aimed at

optimizing the subsequent separation of carbonate from apatite to allow extraction of other

commodities of potential economic interest in this second process scenario are ongoing. The

objective of these ongoing tests is to provide sufficient detail regarding the various potential

extraction processes for the efficiency and cost effectiveness of each to be established in a

robust manner.



3 1B1BINTRODUCTION

3.1 27B27BScope of Work

In November 2011, Coffey Mining Pty Ltd (‘Coffey Mining’) was requested by U3O8 Corp. to

conduct a resource estimate for the Berlin deposit and prepare an Independent Technical

Report (‘ITR’).

This report covers the Resource for the Berlin prospect and is intended to comply with

disclosure and reporting requirements set forth in the Toronto Stock Exchange Manual,

National Instrument 43-101, Companion Policy 43-101CP, and Form 43-101F1.

This report complies with Canadian National Instrument 43-101, for the ‘Standards of Disclosure

for Mineral Projects’ of June 2011 (the Instrument) and the resource and reserve classifications

adopted by CIM Council in November 2004. The report is also consistent with the ‘Australasian

Code for Reporting of Exploration Results, Mineral Resources and Ore Reserves’ of December

2004 (the Code) as prepared by the Joint Ore Reserves Committee of The Australasian Institute

of Mining and Metallurgy, Australian Institute of Geoscientists and Mineral Council of Australia

(JORC).

3.2 28B28BPrincipal Sources of Information

Information used in this report has been gathered from a variety of sources including:

Field observations and reports gathered during the June 2011 field trip.

Supplied U3O8 Corp. technical reports and correspondence.

Various published technical papers.

Geological reports compiled by Dr Richard Spencer from U3O8 Corp.

The principal sources of information used to compile this report comprise digital and hardcopy

excerpts of reports by previous explorers, published scientific papers, technical reports from

U3O8 Corp., and digital exploration and resource modelling data supplied by U3O8 Corp.,

and some published information relevant to the project area and the region in general.

A full listing of the principal sources of information is included in Section 21 of this document.



3.3 29B29BParticipants

The primary author of the report is Mr Neil Inwood – Principle Resource Geologist of Coffey

Mining. Mr Inwood supervised or prepared all sections of the report apart from those relating

to Metallurgy. Mr Inwood takes responsibility for the purposes of the ITR for all sections apart

from Sections 6.4 and 13 and the associated text in the Summary and Conclusions.

Mr John Goode of J.R. Goode and Associates provided oversight for the metallurgical testwork

undertaken by SGS Mineral Services, Lakefield Site, Ontario, Canada.

Dr Paul Miller, Managing Director of Sulphide Resource Processing Pty Ltd provided oversight

for the metallurgical testwork undertaken by SGS Lakefield OreTest in its Malaga, Western

Australia, site.

Mr Goode and Dr Miller take responsibility for the Metallurgy Section of this report

(Section 13) as further stipulated therein (Section 13.2) and the associated text in the

Summary, Conclusions and Recommendations (Sections 1, 25 and 26 respectively).

3.4 30B30BSite Visit

A site visit to the Berlin Project was undertaken by Mr Inwood of Coffey Mining in conjunction

with Dr Richard Spencer and Mr Gabriel Bastias of U3O8 Corp. between 13th and 16th of June

2011. During this visit, Mr Inwood reviewed the data collection procedures, sampling

practices and geology of the project.

Neither Dr Miller nor Mr Goode have visited the site.

3.5 31B31BQualifications and Experience

Coffey Mining is an integrated Australian-based consulting firm, which has been providing

services and advice to the international mineral industry and financial institutions since 1987.

In September 2006, Coffey International Limited acquired RSG Global. Coffey International

Limited is a highly respected Australian-based international consulting firm specialising in the

areas of geotechnical engineering, hydrogeology, hydrology, tailings disposal, environmental

science and social and physical infrastructure.

The primary author of this report is Mr Neil Inwood, a professional geologist with 18 years’

experience in mining and resource geology in Australia, Canada, USA, Europe and Asia.

Mr Inwood is a Fellow of the Australasian Institute of Mining and Metallurgy (‘AusIMM’) and

has the appropriate relevant qualifications, experience and independence to be considered a

Qualified Person as defined in the Canadian National Instrument 43-101.

The initial metallurgical testwork, described in Section 13, was carried out by SGS Mineral

Services, Lakefield Site, in Ontario, Canada, under the guidance of Mr John Goode, a

consulting metallurgist with 49 years experience in metallurgy and related testwork. Mr Goode

is a Fellow of the AusIMM and a member of the Canadian Institute of Mining, Metallurgy and

Petroleum (CIM) and has the appropriate relevant qualifications, experience and independence

to be considered a Qualified Person as defined in the Canadian National Instrument 43-101.



Metallurgical testwork undertaken by SGS OreTest in Perth, Australia, was carried out under

the guidance of Dr Paul Miller who is a metallurgist specialised in hydrometallurgy and has

over 30 years’ experience in the commercial application of processes for the treatment of

sulphide-bearing ore. Dr Miller has a doctorate in Chemical Engineering, is a member of the

Institute of Mining and Metallurgy, London, and is also a Chartered Engineer. He is currently

Managing Director of Sulphide Resource Processing Pty Ltd.

3.6 32B32BIndependence

None of the associated authors of this report have any material interest in U3O8 Corp. or

related entities or interests. Their relationship with U3O8 Corp. is solely one of professional

association between client and independent consultant. This report is prepared in return for

fees based upon agreed commercial rates and the payment of these fees is in no way

contingent on the results of this report.

3.7 33B33BAbbreviations

A listing of abbreviations used in this report is provided in Table 2.7_1.



Table 2.7_1

List of Abbreviations

“ inches HRD half relative difference P2O5 Diphosphorus Pentoxide

µm microns ICP-AES inductivity coupled plasma atomic emission spectroscopy P80 80% of sample passing

3D three dimensional ICP-MS inductivity coupled plasma mass spectroscopy ppb parts per billion

AAS atomic absorption spectrometry IRR Internal Rate of Return ppm parts per million

Ag Silver ISO International Standards Organisation ppt. precipitate

Al Auminium JORC Joint Ore Reserves Committee PQ size of diamond drill rod/bit/core

AusIMM Australasian Institute of Mining and Metallurgy K Potassium Q2 second quarter

BSE Backscattered electron microscope kg kilogram QAQC Quality assurance, quality control

C Carbon km kilometres QC quality control

C$ Canadian dollars km² square kilometres QQ quantile-quantile

Ca Calcium ktonnes thousand tonnes Rb Rubidium

CC correlation coefficient IAN Instituto de Asuntos Nucleares (Colombia) RC rock chip

CIM Canadian Institute of Mining, Metallurgy & Petroleum lb pounds Re Rhenium

cm centimetre ID² Inverse Distance squared REE Rare earth element

CNI Canadian National Instrument m metres RL (Z) reduced level

Co cobalt M million RQD rock quality designation

COG Cutoff Grade M Mole/molar SEM Scanning electron microscope

cps Counts per second m³/t cubic metres per tonne SD standard deviation

CRM certified reference material or certified standard Ma Million years SG Specific gravity

Cu copper m.a.s.l. Metres above sea level Si silica

CV coefficient of variation Mg Magnesium Sr Strontium

DEM Digital evelvation model mg/L milligrammes per litre t tonnes

DDB diamond drillhole Mkg Million kilograms t/m³ tonnes per cubic metre

DGPS Differential Global Positioning System ml millilitre Ta Tantalum

DMS Dense Media Separation Mlb Million pounds TB Trench prefix

DTM digital terrain model mm millimetres Th Thorium

E (X) easting Mn Manganese tpa tonnes per annum

eU3O8 equivalent U3O8 Mo Molybdenum TSX Toronto Stock Exchange

Fe iron Mt Million Tonnes U Uranium

ft. foot Mtpa million tonnes per annum U3O8 tri uranium octoxide

g gram N (Y) northing UNDP United Nations Development Program

G&A General and Administration Nb Niobium UO2 Uranium dioxide

g/L grammes per litre Nd Neodymium UO4 Uranium tetroxide

g/m³ grams per cubic metre Ni Nickel UTM Universal Transverse Mercator

g/t grams per tonne NORM Naturally occurring radioactive material V Vanadium

GDP Gross Domestic Product NPV V2O5 Vanadium pentoxide

GPS Geographical positioning system NQ net present value WGS84_18n World Geodetic System 1984 - zone 18 north

GRS Gamma Ray Spectrometer NQ2 size of diamond drill rod/bit/core Y Yttrium

HARD half the absolute relative difference ºC degrees centigrade Y2O3 Yttrium oxide

HQ size of diamond drill rod/bit/core OK Ordinary Kriging XRD X-Ray diffraction

hr hours P Phosphorus XRF X-Ray fluorescence


Berlin Project, Colombia – MINEWPER00790AC Page 7 National Instrument NI 43-101 Report – 2 March 2012

4 2B2BRELIANCE ON OTHER EXPERTS

The information regarding title and the legal status on the exploration concessions that

constitute the Berlin Project, as described in Section 4 below, was provided by Dr Hernán

Rodríguez, independent Colombian counsel to U3O8 Corp. Dr Rodríguez is based in Bogota,

Colombia and is a partner with Norton Rose Group, an international law firm with offices across

Asia, Europe, Canada and Latin America. Dr Rodríguez has been practicing law since 1995,

and has participated in hydrocarbon, mining and infrastructure projects in Colombia, including

projects for ports, railways and oil refineries, and has been actively involved in mergers and

acquisitions transactions, especially in the mining industry. Coffey Mining is not qualified to

independently verify this data.

The details regarding environmental legislation and requirements are based upon information

supplied to Coffey Mining from U3O8 Corp. in various documents, and as part of due

diligence documentation. Coffey Mining is not qualified to independently verify this data.

The documents referred to include:

The title opinion dated February 8, 2012 prepared by Dr Rodríguez of Norton Rose

(Rodríguez, 2012);

Copy of the name change certificate from Energentia Ltd. to Gaia Energy Investments Ltd;

Copies of the concession contracts that comprise the Berlin Project; and

Summary Report of the Environmental Activity and Social Programs of the Berlin Project

(Serna, 2011).



5 3B3BPROPERTY DESCRIPTION AND LOCATION

5.1 34B34BRegulatory Framework for Mineral Properties in Colombia

5.1.1 85B85BOverview

The Colombia Government has actively created an environment that is friendly to responsible

mining. Although mining accounted for only 1.5% of GDP between 2003 and 2008, 21% of

direct foreign investment was made in the mining and exploration industry in that period

(National Department of Statistics (DANE)). Early indications are that this investment is

leading to new discoveries that will boost the mining industry’s contribution to GDP. Between

2004 and 2008, the mining industry contributed 20% of Colombia’s exports.

5.1.2 86B86BLegal and Regulatory Framework of Mineral Concessions

In Colombia, exploration and exploitation of mining resources, such as uranium, are

formalised by execution of a concession contract (the “Concession Contract”) with the

correspondent national mining authority pursuant to the mining legislation Law N° 685/01, duly

amended by Law 1382 of 2010.

Since 1940, the mining authority has been the Ministry of Mines and Energy. The Ministry

has, in turn, delegated some mining-related matters to national and provincial authorities (the

‘Mining Authority’). Specifically, the National Institute of Geology and Mining (INGEOMINAS)

is responsible for managing royalties and maintaining the national register of Concession

Contracts. Specific Provincial Governments are charged with the granting, execution and

performance of Concession Contracts and other related administrative proceedings within

their respective provincial boundaries. This is the case for the Caldas Province, in which the

Berlin Project is located, whereby the Province manages all exploration and mining-related

activities for minerals found within the province, except for coal and emeralds, which are

managed by INGEOMINAS.

By means of Decree 4134 of November 3rd, 2011, the Colombian Government created the

National Minerals Agency (the ‘Agency’). As per that Decree, the Agency will assume

responsibility for the granting, execution and administration of Concession Contracts

throughout Colombia. The Agency is scheduled to assume that responsibility in May, 2012.

Under Colombian law, foreign individuals and corporations have the same rights as

Colombian individuals and corporations. Foreign companies are required to constitute a

branch, subsidiary or affiliate in Colombia before they may be granted a Concession Contract.

U3O8 Corp’s wholly-owned subsidiary, Gaia Energy Investments Ltd., has a branch in

Colombia called Gaia Energy (Colombia) Ltd., which has Concession Contracts covering the

Berlin Project area.



5.1.3 87B87BRepatriation of Funds and Payment of Dividends

Companies that have sales in foreign currencies are required to repatriate these amounts in

Colombian Pesos through authorised foreign exchange intermediaries. However, through

Article 16 of Law 9 of 1991 and Decree 2058 of 1991, branches of foreign companies

undertaking exploration and exploitation of uranium, petroleum, natural gas, coal or

ferronickel, are exempt from this repatriation obligation. Instead, branches of companies

involved in the exploration and mining of these commodities are required to repatriate the

amounts necessary to pay expenses in local currency.

5.1.4 88B88BCanada-Colombia Free Trade Agreement

Colombia and Canada signed a free trade agreement in June 2010.

5.1.5 89B89BMineral Concessions

The Concession Contracts for the Berlin Project were executed and registered prior to

Law 1382 of 2010 coming into effect. The Berlin Concession Contracts are valid for a 30 year

term that can be extended for 30 additional years. The initial term of the Concession

Contracts comprises the following three phases:

Exploration - three years with possible extension for eight additional years.

Construction - three years for the construction and assembly of the infrastructure,

extendable for one additional year; and

Exploitation - the remaining years for the exploitation stage, extendable for a further

30 years upon request by the concession holder.

Concession Contracts granted after February 9, 2010, when Law 1382 of 2010 came into

effect, have the same initial 30 year term, but are extendable for 20 additional years. In

addition, the five year term for the exploration phase may be extended for a total of 11 years

prior to the construction phase.

Concession Contracts for exploration convey the right to explore the defined areas for

specified metals or minerals. A concession owner has the first right to include additional

commodities and metals to the original Concession Contract. The rights of the Concession

Contract can be assigned totally or partially to another party, subject to prior notice and

authorization by the Mining Authority and as long as the obligations under the Concession

Contract have been duly complied with.

Surface rights are separate from the exploration or mining rights. None of the Concession

Contracts covering the Berlin Project imply any surface rights – acquisition of surface rights

must be negotiated directly with the landowners. The Concession Contracts are renewed

annually, provided that work commitments and property payments due to the Mining Authority

have been met, such as those shown below (Item 4.6) for each concession on the Berlin

Project.



30 days prior to the expiration of the exploration period, the concession owner is required to

file and obtain approval of a Working and Construction Plan from the Mining Authority as well

as an Environmental Impact Assessment (EIA) from the relevant environmental authority in

order to advance to the construction and assembly stage.

5.1.6 90B90BOther Required Permits and Environmental Liabilities

Environmental Mining Insurance

Within 10 days following the execution of the Concession Contract, an environmental mining

insurance policy must be obtained by the concession holder to guarantee compliance with

mining and environmental obligations, the payment of fines and the unilateral termination of

the agreement by the Mining Authority. The insured value is calculated as follows for the

different stages:

Exploration – 5% of annual estimated work expenditures.

Construction – 5% of the annual investment towards mine construction.

Exploitation – 10% of the result of multiplying the estimated annual production by the

price of the mineral being extracted, as determined by the Government.

The insurance policy must be in full force and effect throughout the life of the Concession

Contract and is renewed annually as part of the fulfilment of obligations to maintain the

Concession Contract in good standing.

Environmental License

No specific environmental license is required for the exploration stage. However, all work

must be done in accordance with environmental guidelines issued by the Ministry of Mines

and Energy and the Ministry of the Environment. Three drill-related permits are required for:

solid waste management;

water use; and

permission to establish drill pads.

The water-use permit typically takes about six months to process while the other drill-related

permits typically take 2-3 months. Water from one permitted site can be used for various drill

platforms in the vicinity.

In order to commence mine construction, an EIA must be completed in order to obtain an

environmental licence from the respective environmental authority.

No environmental liabilities on the Berlin Project are known to the author at this time. There

are no other known significant factors and risks that may affect access, title, or the right or

ability to perform work on the Berlin Project.



5.1.7 91B91BConcession Fees

During the stages of exploration, construction and assembly, the concession holder must pay

an annual surface fee to the Mining Authority based on the assigned areas and the number of

years of the exploration period (see Table 4.1.7_1 for fees paid on Concession Contracts

granted prior to February 9, 2010). The basis of the fee is the Colombian minimum daily

wage, which is approximately US$10.00 at present.

Table 4.1.7_1

Berlin Project

Table of Annual Concession Fees

Concession Size Fees per Hectare (“Ha”)

Approximately

Up to 2,000Ha US$10.00

2,000Ha – 5,000Ha US$20.00

5,000Ha – 10,000Ha US$30.00

Concession Contracts granted after February 9th, 2010 are subject to a different annual fee of

approximately US$10.00 per hectare for the first five years of exploration, increasing by 25%

of the annual wage per hectare for each year thereafter. During the construction phase, the

annual fee is frozen at the maximum level paid in the last year of the exploration phase.

5.2 35B35BRoyalties

During the exploitation stage, the concession holder is required to pay a net smelter royalty

(“NSR”) to the Mining Authority. The rate of the royalty differs by commodity (Table 4.2_1). In

addition, a 2% NSR is payable to AngloGold Ashanti on commencement of commercial

uranium production from Concessions C (No 664-17), D (No 736-17) and E (No IFM 08221X).

This royalty derives from the acquisition of the properties by KPS Ventures Ltd. from

AngloGold Ashanti’s subsidiary, Sociedad Kedahda SA, as described in Section 6.1.

Table 4.2_1

Berlin Project

List of NSR on the Mining of Various Commodities

Commodity NSR (%)

Uranium 10

Vanadium, Phosphate, Molybdenum, Yttrium, Rhenium, Iron, Copper, Platinum 5

Gold, Silver 4

Nickel 12

Coal 5-10

Salt 12

Construction Materials 1

5.3 36B36BArea of the Berlin Project Properties

The Berlin Project lies within an area of 10,681 Hectares covered by five contiguous

concessions in the Province of Caldas in central Colombia.



5.4 37B37BLocation of the concessions

Details of the Gauss-Kruger and UTM co-ordinates of the corner points of the five concessions

that cover the Berlin Project are provided in Table 4.4_1.

Table 4.4_1

Berlin Project

Corner Points of the Berlin Project Concessions (Gauss-Kruger and UTM, Zone 18 North Coordinates)

Concession Corner Point

Gauss- Kruger UTM

Northing Easting Northing Easting

755-17

1 1111460 902000 619371 504207

2 1111920 900210 619828 502417

3 1110000 899180 617907 501391

4 1110000 897000 617904 499212

5 1115000 897000 622901 499204

6 1113000 902000 620910 504205

7 1115000 902070 622909 504272

8 1113000 902070 620910 504275

756-17

1 1107000 897830 615220 499671

2 1107000 897000 615219 498842

3 1110000 897000 618217 498837

4 1110000 899230 618221 501066

664-17

1 1115310 907212 623540 509035

2 1115310 904000 623535 505835

3 1115150 904000 623375 505835

4 1115150 902060 623372 503886

5 1115000 902060 623223 503886

6 1115000 902070 623223 503896

7 1113000 902070 621224 503899

8 1113000 902000 621223 503830

9 1111460 902000 619684 503832

10 1111920 900210 620141 502042

11 1110000 899180 618221 501016

12 1110000 899230 618221 501066

13 1107000 897830 615220 499671

14 1107000 897000 615219 498842

15 1110000 897000 618217 498837

16 1115000 897000 623215 498829

17 1115000 899110 623218 500938

18 1115310 899110 623528 500937

19 1115310 895513 623522 497343

20 1106763 895513 614980 497356

21 1106763 907212 614998 509049

IFM 08221X

1 1115310 899109 623215 501312

2 1115310 900001 623216 502203

3 1115310 901266 623218 503467

4 1115000 901266 622908 503468

5 1115000 899110 622905 501313

736-17

1 1115310 901000 623218 503202

2 1115310 895513 623209 497718

3 1119500 895513 627397 497711

4 1120000 896500 627898 498697

5 1120000 901722 627906 503916

6 1117840 901722 625746 503920

7 1117840 901000 625746 503198

8 1115310 901000 623218 503202

These co-ordinates are measured in metres.



5.5 38B38BLocation of Mineralised Zones Relative to the Properties

The location of the mineralised unit in the Berlin Project is shown relative to the concession

boundaries in Figure 4.5_1.

Figure 4.5_1

Location of the Berlin Project Exploration Concessions

* Contracts 736-17; IFM-08221X_N to be assigned to Gaia Energy

The concessions are overlain on a digital elevation model. The green outline shows the location of the target Cretaceous sedimentary sequence for reference.

5.6 39B39BDetails Pertaining to the Concessions

5.6.1 92B92BConcession 1 (File No 755-17)

Legal Status

Concession Contract:

Executed on 23th October 2007;

Registered on 9th November 2007;

Original expiry: 9th November, 2010

Extension Application:

Application has been made for the two-year extension for the concession - expiry:

9th November 2012.



Mineral concession phase:

Exploration

Size:

2,122.61 Hectares

Location:

Municipality of Samana, Department of Caldas - Berlin area.

Commodity

Uranium and radioactive elements;

Vanadium, molybdenum and phosphate;

Any other mineral that can be economically exploited in the concession area.

Current Concession Owner:

Gaia Energy (Colombia) Ltd

Annual Surface Fee:

USD 43,000

Environmental License:

Not applicable during the exploration stage.


Legal Status


Executed on 23rd October 2007;

Registered on 9th November 2007;

Original expiry: 9th November, 2010.


application has been made for the two-year extension for the concession - expiry:

9th November, 2012.


Exploration

Size:

459 Hectares

Location:




Commodity



Any other element of economic value that can be extracted from the concession.


Gaia Energy (Colombia) Ltd

Annual Surface Fee:

USD4,700




Legal Status



Registered on 7th December 2007.

Original expiry: 7th December 2010.


Application has been made for two-year extension for the concession - expiry:

7th December 2012.


Exploration

Size:

7,304.9 Hectares

Location:

Municipality of Samana, Department of Caldas – Berlin area.

Commodity

Gold;






Current Concession Owner

Gaia Energy (Colombia) Ltd;

The Mining Authority issued Resolution 0241, dated 1st February 2010, authorizing the

assignment of the Concession Contract from AngloGold Ashanti to Energentia.

Annual Surface Fee:

USD223,600



5.6.4 95B95BConcession 4 (File No IFM 08221X)

Legal Status


Executed on 12th December 2007;

Registered on 21st December 2007;

Original expiry: 21st December 2010.


Application has been made for the two-year extension for the concession – expiry:

21st December 2012.


Exploration

Size:

74.5 Hectares

Location:

Municipality of Samana, Department of Caldas – Berlin area.

Commodity

Gold, Silver, copper, zinc, platinum, molybdenum;


Vanadium and phosphate;

Any other mineral that can be economically exploited from the concession area.



Current Concession Owner

Gaia Energy (Colombia) Ltd.

The Mining Authority issued resolution 3837 dated 22nd June 2010 allowing the transfer

of the concession from AngloGold Ashanti to Energentia. The resolution has been filed

with the Mining Registry.

Annual Surface Fee:

Approx. USD760




Legal Status



Registered on 9th November 2007.

Original expiry: 9th November, 2010.


Application has been made for two-year extension for the concession - expiry:

9th November 2012.


Exploration

Size:

2,704 Hectares

Location:


Commodity

Gold, silver, zinc, platinum, molybdenum


Vanadium and phosphate;





AngloGold Ashanti. The assignment of this concession to Energentia by AngloGold

Ashanti, was filed with the Mining Authority on 14th January, 2010. The Resolution

approving assignment has been drafted and is awaiting execution by the Governor

of Caldas Province.

Annual Surface Fee:

Approx. USD 55,000



5.7 40B40BRelated Agreements

As described above, through the acquisition of Concessions No 664-17, No 736-17 and

No IFM 08221X by KPS Ventures Ltd. from AngloGold Ashanti’s subsidiary, Sociedad

Kedahda SA, AngloGold Ashanti maintains a 2% NSR on uranium production on commencement

of commercial uranium production. A payment of approximately US$250,000 is due to

AngloGold Ashanti on assignment of the last of the three concessions (Concession 736-17) that

originally belonged to AngloGold Ashanti.

5.8 41B41BComments

Security factors may influence future operational decisions however the author understands

that regionally the security situation is improving. Coffey Mining is not aware of any other

significant factors which would affect the access, title or the right and ability of the Company to

perform work on the property.



6 4B4BACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

6.1 42B42BTopography, Elevation and Vegetation

The Berlin Project area lies on the eastern flank of the Cordillera Central at approximately

5° 31’ North and 75° West (Figure 5.1_1). The Berlin area is characterised by steep hills

separated by perennial streams at an altitude of 400-1,500 metres above mean sea level

(Figure 5.1_2). The Project lies within the catchment basin of the Rio Manso, which flows

eastward into the Rio Negro and then into the Rio Magdalena which flows northwards

between the Cordillera Central and Cordillera Occidental, discharging into the Caribbean Sea.

Figure 5.1_1

General Location of the Berlin Concession Areas in Caldas Province, Colombia



Figure 5.1_2 Photo showing the Typical Landscape of the Berlin Project

6.2 43B43BAccess and Infrastructure

The Berlin Project area is located in Caldas province of Colombia some 80km northeast of the

provincial capital, Manizales. It lies approximately 150km northeast of the national capital,

Bogota, and 100km southeast of the city of Medellin which also has an international airport.

Road access is good with the project area lying 59km from the paved Medellin - Bogota

highway. A secondary paved road leads 50km west from the Bogota-Medellin to the village of

Norcasia. The road continues, unpaved, 9km beyond to Berlin village. Agriculture is the

principal economic activity in the area.

The Magdalena River is navigable by barge from the town of Puerto Boyoca, some 65km

northeast of the project area to the port of Barranquilla on the Caribbean coast.

A railway line leads from the town of El Dorado on the Rio Magdalena to the port town of

Santa Marta on the Caribbean coast. Although the railway line is not currently in use, the

government has flagged it as a priority infrastructure project for completion by about 2015.

The 395MW La Miel hydroelectric dam is located approximately 12km from the central part of

the project area (Figure 5.2_1).



Figure 5.2_1 General Location of the Berlin Project

Google satellite image showing secondary road access from the town of La Dorada to the villages of Norcasia and Berlin

6.3 44B44BClimate

The Berlin Project is within a tropical area with average annual temperatures of 23° Celsius and

average rainfall of 1,000mm. Rainfall occurs throughout the year with a peak in January to

March.



7 5B5BHISTORY

7.1 45B45BPrior Ownership of the Berlin Property

Minatome, a French exploration company, which has now been incorporated into AREVA,

undertook the most detailed exploration of the Berlin Project between 1979 and 1981. Its

work involved basic geological investigation followed by the core drilling, initial metallurgical

testwork and estimation of a uranium resource (not NI 43-101 compliant). Minatome is

reported to have withdrawn from the Berlin area when the company was nationalised by the

French government in 1981, which coincided with a slump in uranium prices.

After Minatome’s withdrawal from the project, the concessions reverted to the State. The

United Nations Development Program (UNDP) reviewed the technical work undertaken on the

project in 1982 and focused on the potential to recover uranium, molybdenum, vanadium and

phosphate from the Berlin area. The UNDP suggested that Dense Media Separation may

play a part in recovery of the various commodities.

Energentia Ltd, formerly KPS Ventures Ltd., applied for exploration concessions directly from

the State and also entered into an agreement with AngloGold Ashanti to acquire properties in

the Berlin Project area in 2007. Specifically, Energentia entered into an agreement with a

wholly owned subsidiary of AngloGold Ashanti, Sociedad Kedahda SA. Due to the slow

nature of related legal procedures, one of these property transfers is still in process and

completion is pending as described in Section 4 above. Subsequently, Mega Uranium Ltd

purchased Energentia Ltd on 1st May, 2008.

U3O8 Corp. closed the acquisition of Mega’s South American assets, including the Berlin Project,

on May 7th 2010. On 26th November 2010 Energentia Ltd name was changed to Gaia Energy

Investments Ltd and the Colombian branch name to Gaia Energy (Colombia) Ltd.

7.2 46B46BPrior Exploration

Phosphate was identified in the lower parts of the Cretaceous sedimentary sequence in the

Berlin area in 1968 (SRK, 2006). Uranium was identified in the phosphatic black shales in a

regional radiometric prospecting program undertaken by the Colombian Instituto de Asuntos

Nucleares (IAN) between 1977 and 1983. Minatome obtained permission from IAN to explore

the Berlin Project area for uranium in 1979. Field-based exploration carried out by Minatome

identified the black shale near the base of the Cretaceous sequence in the Berlin area as having

significant uranium grades. Rock-chip sampling resulted in the identification of highly

anomalous uranium values over the entire strike length of the synform in the Cretaceous

sequence in the Berlin area (Figures 6.2_1 and 6.2_2).

Minatome’s assay results from surface rock-chip sampling were substantiated by independent

sampling undertaken by Naranjo (1983), at the Universidad National in Bogota, who reported

mapping the uraniferous black shale over a strike length of approximately 3,000m in the

southern part of the synform in the Cretaceous sequence. Analyses from seven channel rock

chip samples and one point sample (Sample 6) were reported as shown in Table 6.2_1.



Figure 6.2_1 Channel Sample Uranium Grades in Uraniferous Black Shale, Northern Berlin Project Area

Geological map of uranium grades of channel samples in the uraniferous black shale. Red figures are uranium grades in ppm and the length of the channel samples is shown in centimetres in black.



Figure 6.2_2 Channel Sample Uranium Grades in Uraniferous Black Shale, Southern Berlin Project Area

Geological map of uranium grades of channel samples in the uraniferous black shale. Red figures are uranium grades in ppm and the length of the channel samples is shown in centimetres in black.



Table 6.2_1

Berlin Project

Rock-Chip Sample Assay Results from Radioactive Unit, Southern Berlin Project Area

Sample No.

Coordinate Height m.a.s.l.

Radiometry cps (SPP2)

Thickness (m)

U3O8 %

P2O5 %

Mo %

V2O5 % UTM N UTM E

1 618341 503684 720 9,000 2.60 0.544 7.14 n/a n/a 2 619475 503482 680 6,000 1.50 1.305 10.50 n/a n/a 3 619060 503188 690 6,000 2.00 0.258 (b) 10.15 n/a n/a 4 618161 503469 820 6,000 2.00 0.174 (b) n/a n/a n/a 5 617655 512386 860 15,000 1.00 0.985 n/a n/a n/a 6 617561 503690 840 15,000 Punctual 0.836 (c) n/a n/a n/a 7 617411 503710 940 15,000 1.50 0.495 n/a n/a n/a

8 (a) 623508 503401 810 6,000 2.00 0.648 11.10 0.49 1.15

Co-ordinates are given in UTM zone 18 north measured in metres. (Cps = radiometric counts per second and n/a = not analysed). Source: Naranjo (1983)

Table 6.2_2

Berlin Project

Rock-Chip Sample U3O8 Assay Values and Thickness

Sample No Location - Coordinates U3O8 Grade

(%) Sample Thickness

(metres) UTM North UTM East

BW-375 619725 502610 0.197 0.2 6W-200 618132 503055 0.214 1.4 BW-175 617962 503209 0.310 1.8 BW-150 617805 503231 0.239 2.3 BW-125 617494 503367 0.184 1.1 BW-100 617337 503543 0.144 1.0 BW-75 617174 503673 0.177 1.8 BW-50 616953 503800 0.175 0.7 BW-25 616744 503913 0.184 2.4

BE-0 616625 504230 0.196 1.0 BE-01 616731 504213 0.293 2.0 BE-25 616843 504071 0.157 1.1 BE-50 617034 503930 0.124 0.9 BE-75 617274 503899 0.163 2.6 BE-100 617523 503739 0.269 1.2

BE-125 617675 503527 0.314 0.9 BE-150 617886 503443 0.317 2.3 BE-175 618108 503379 0.194 1.1 BE-200 618362 503346 0.127 1.3 BE-225 618654 503237 0.088 0.7 BE-250 618891 503250 0.088 1.1 BE-275 619095 503215 0.106 0.4 BE-300 619304 503381 0.102 1.4 BE-325 619487 503499 0.384 1.5 BE-400 620086 503322 0.157 0.2 BE-450 620647 503319 0.328 1.3

BE-475 620939 503568 0.366 1.1 BE-500 621162 503508 0.098 0.9 BE-525 621419 503532 0.148 1.3 BE-675 622910 503509 0.315 0.4 BE-650 622622 503618 0.241 2.5 BE-875 624966 503185 0.097 1.3

BE-900 625226 503131 0.155 1.2 BE-925 625480 503117 0.199 0.0 BE-950 625731 503109 0.148 0.2 BE-1025 626344 503243 0.096 0.2

Coordinates are in UTM zone 18 north measured in metres. The location of these samples is shown in figures 6.2_1 and 6.2_2. Source: (Castaño, 1981)



Minatome’s exploration concentrated on the southern 5km of the 10.5km long syncline where

access is easier and where outcrop of the sedimentary sequence is generally better in

comparison to the north. The apparent consistency of uranium grades along strike led

Minatome to excavate 20 trenches and 3 adits, the latter with the objective of confirming

mineralisation in fresh exposures beneath the saprolite. The location of the adits is shown in

Figure 6.2_3 and the assay results reported by Minatome are shown in Table 6.2_3.

Figure 6.2_3 Location of the Adit Portals Excavated by Minatome During 1979-1981 Exploration Program,

Southern Berlin Project



Table 6.2_3

Summary Assay Results of Channel Samples taken through the Uraniferous Unit in the Three Adits excavated by Minatome and whose Locations are shown in Figure 6.2_3

Tunnel 1 Tunnel 2 Tunnel 3

Length (m) 48 24 40 Thickness (m) 1.8 3.75 3.25 U3O8 (ppm) 362.2 1,099.76 587.6 V (ppm) 3490 11,069 10,966 Mo (ppm) 406 2,306 175 P2O5 (%) 6.51 4.7 8.9

Minatome then drilled 11 widely-spaced drillholes for a total of 2,136m in 1980. Although six

of these drillholes are reported to have reached the target depth, nine are reported to have

intersected anomalous uranium values. IAN is reported to have drilled six drillholes in the

Berlin Project area in 1982 and 1983 and three of these holes are reported to have reached

the target horizon and to have intersected grades similar to those reported by Minatome over

similar true widths (SRK, 2006; Figure 6.2_4).

Figure 6.2_4

Historical Drilling in the Berlin Area

Only drilling undertaken by U3O8 Corp. has been used in the Resource estimate detailed in

Section 14.



7.3 47B47BHistoric Resource Estimates

7.3.1 97B97BHistoric Resource Estimate

Minatome undertook a provisional, non CIM-compliant, grade estimate, using polygonal

methods based on assays of mineralised intervals of the drillholes whose collar positions are

shown in Figure 6.2_4 and rock-chip channel samples from the adits (Table 6.2_2) in 1981. The

historic estimate by Minatome, which covered the southern 4.5km of the 10.5km-long

mineralised trend in the Berlin area, was 12.9 million tonnes at a grade of 0.13% U3O8 for a total

of approximately 38 million pounds of contained U3O8 (Castaño, 1981). Minatome’s estimate

was not done in accordance with National Instrument 43-101 and therefore should not be

construed as a Resource or an indication of the Resource endowment of the project. The

estimate has been included only for historical context of the project.

7.4 48B48BMetallurgical Testwork

Minatome (Nancy, France) conducted various metallurgical tests on the black shales from the

Berlin Project in 1979 (Roussemet & Houot, 1979). Simple acid leaching resulted in a 75%

recovery of uranium, but with the consumption of 130kg of acid per tonne of uraniferous shale.

Provisional metallurgical testwork on a 35kg rock-chip sample from the adits showed the

following distribution of uranium (Roussemet & Houot, 1979):

5-10% of the uranium occurs on the surface of coarse fragments and is liberated during

crushing of the host-rock

55-60% of the uranium is associated with phosphate in the 40-200 micron (“µm”) fraction.

Testwork shows that the phosphate is amenable to flotation; and

Approximately 30% of the contained uranium occurs with the fine fraction, suspected to

be adsorbed onto illite and other clays, and was liberated on ultrafine grinding to a

nominal grain size of eight microns.

The conclusion from Minatome’s metallurgical testwork was that uranium recovery was

approximately 85% using a combination of flotation and ultrafine grinding of the fine fraction

(Roussemet & Houot, 1979). This work did not include an estimate of cost of processing.



8 6B6BGEOLOGICAL SETTING AND MINERALISATION

8.1 49B49BRegional Geology

Basement in the Berlin area is composed of greenschist to lower amphibolite grade pelitic and

graphite-bearing schists that are correlated with the Cajamarca – Valdivia terrane (CA-VA).

Elsewhere the CA-VA contains rocks of ophiolitic origin such as olivine gabbro, pyroxenite,

chromitite and serpentenite. Geochemical signatures suggest that these intrusive rocks are of

intraoceanic arc and continental margin affinity. They form part of a parautochthonous

sequence that was accreted onto the Chicamocha terrane along the Palestina fault during the

Ordovician to Silurian (Figures 7.1_1 and 7.1_2). Further south, the CA-VA was accreted

directly on to the Guyana Shield (Cediel et al. 2003).

The sedimentary sequence that contains the mineralisation at Berlin is lower Cretaceous in

age and is part of an extensive marine basin, the Magdelena Basin that was formed as a

thermal sag phase over active rifting involving extensive half-grabens of Jurassic to lower

Cretaceous age. Inversion of the extensional basin started in the Paleogene and accelerated

in the Late Miocene – Pliocene. This basin inversion resulted in the formation of the

Cordillera Oriental (Eastern Cordillera) in Colombia (Figure 7.1_2).

The Berlin Project lies on the eastern flank of the Cordillera Central where remnants of the

Cretaceous marine sequence overlie schists of the CA-VA (Figure 7.1_1).

Several batholiths occur in the vicinity of the Berlin Project (Figure 7.1_3). The Sansón

Batholith, located some 20km west of the project area, is a metaluminous I-type calc-alkaline

body of Triassic to Jurassic age that intruded the CA-VA. The Antioquia Batholith lies about the

same distance to the north of the project area and is middle to late Cretaceous in age with age

dates of 90-58Ma having been obtained (Cediel et al., 2003). It has a similar metaluminous, I-

type, calc-alkaline composition to the Sansón Batholith. The Samaná Batholith is located

immediately to the west of the Cretaceous sedimentary sequence at Berlin, and some igneous

phases intrude the sedimentary sequence that hosts the mineralisation.

Folding of the lower Cretaceous sedimentary sequence at Berlin is assumed to have taken

place in response to inversion of the basin which started in the Paleogene. The age of the

associated axial planar cleavage to folds in the Berlin area provides a time-line against which

the relative age of mineralisation can be established.

8.1.1 98B98BLithology and Stratigraphy

Basement

The Paleozoic basement in the Berlin Project area consists of metamorphic rocks – mainly

greenschists, graphitic schists and quartzites. Nuñez et al. (1979) maintain that the Paleozoic

sequence in the Berlin area may correlate with the Cajamarca Group of Nelson (1957), the

Ayura-Montebello Group of or the Valdivia Group. More recently the sequence has been

referred to as the Metamorphic Rocks of the Central Range (Feininger et al., 1972),

Cajamarca Terrain (Etayo-Serna et al., 1986), Tahamí Terrain (Toussaint and Restrepo,

1988) and the Cajamarca Complex (Maya and González, 1995). Here it is referred to as the

Cajamarca-Valdivia terrane, following the usage by Cediel et al. (2003).



Figure 7.1_1 Main Tectonic Components of Colombia

The inset map shows the location of the section shown in Figure 7.1_2.

GS = Guiana Shield; GA = Garzon massif; SP = Santander massif – Serrania de Perija; ME = Sierra de Merida; SM = Sierra Nevada de Santa Marta; EC = Eastern Cordillera; CO = Carora basin; CR = Cordillera Real; CA-VA = Cajamarca-

Valdivia terrane; sl = San Lucas block; ib = Ibague block; RO = Romeral terrane; DAP = Dagua-Piñon terrane; GOR = Gorgona terrane; CG = Cañas Gordas terrane; BAU = Baudo terrane; PA = Panama terrane; SJ = San Jacinto terrane; SN = Sinu terrane; GU-FA – Guajira-Falcon terrane; CAM – Caribbean Mountain terrane; Rm = Romeral melange; fab = fore

arc basin; ac = accretionary prism; tf = trench fill; pd = piedmonte;

1 = Atrato (Choco) basin; 2 = Tumaco basin; 3 = Manabi basin; 4 = Cauca-Patia basin; 5 = Upper Magdalena basin; 6 = Middle Magdalena basin; 7 = Lower Magdalena basin; 8 = Cesar-Rancheria basin; 9 = Maracaibo basin; 10 = Guajira basin;

11 = Falcon basin; 12 = Guarico basin; 13 = Barinas basin; 14 = Llanos basin; 15 = Putumayo-Napo basin; red dot = Pliocene-Pleistocene volcano.

(Cediel et al, 2003)



Figure 7.1_2 Cross Section Through the Colombian Cordileras Showing the Regional Context of the Berlin Project

Regional context of the Berlin Project area that lies adjacent to the Palestina Fault system (2) on the east flank of the

Cordillera Central (from Cediel et al., 2003).

Principal sutures: 1 = Grenville (Orinoco) Santa Marta – Bucaramanga – Sauza faults; 2 = Ordovician-Silurian Palestina fault system; 3 = Aptian Romeral-Peltetec fault system; 4 = Oligocene-Miocene Garrapatas-Dabeiba fault system; 5 = late Miocene Atrato fault system.

Abbreviations: K-wedge = Cretaceous wedge; CA-VA = Cajamarca-Valdivia terrane; MMB = Middle Magdalena Basin; sl = San Lucas block; (Meta-)Sedimentary rocks: pz = Palaeozoic; K = Cretaceous; P = Paleogene; N = Neogene.

Figure 7.1_3

Regional Geological Setting of the Berlin Project

Berlin Project concessions are shown in yellow.



The main metamorphic event has been dated at Paleozoic to Early Mesozoic (Restrepo et al.,

1991; Maya, 1992; Maya and González, 1995), although other authors suggest a broader

interval, from the Neoproterozoic to the Early Mesozoic. González (1980) and Alvarez (1979)

interpret this basement sequence as being geosynclinal in nature.

The sedimentary sequence that contains the mineralisation in the Berlin Project area is

Cretaceous in age and rests unconformably on the Palaeozoic basement (Figure 7.1_3). The

sequence is made up of a clastic lower unit that fines upwards into a black, organic-rich

mudstone unit. Fossiliferous limestones are locally developed within the black mudstone

sequence and the fossil assemblage indicate a Valanginian – Hauterivian (Early Cretaceous)

age and that the mudstones are marine in origin.

Cretaceous Sedimentary Rocks

Valle Alto Formation

Defined by González (1980), this refers to a sandstone-dominated sequence that lies to the

west of the Berlín Project. Based on molluscs and associated plant fossils, an Early

Cretaceous age was determined (Etayo-Serna, 1985) and this age was confirmed by a more

recent review of the flora (Vakhrameev, 1991). The sedimentary sequence is interpreted to

have formed in a transitional environment that includes both continental and marine facies

that formed during a marine transgression that flooded onto the metamorphic basement of

Cajamarca Complex (Etayo-Serna et al., 2003).

Abejorral Formation

The Abejorral Formation (Bürgl and Radelli, 1962) is a Cretaceous sequence that is preserved

in outliers in the Caldas Province, and in the adjacent province to the north (Antioquia). In some

sectors this formation is discordant over the rocks of the Cajamarca Complex and in faulted

contact with the Valle Alto Formation. Based on ammonites, González (1980) placed this

formation in the Late Aptian – Middle Albian interval. Mapping by González (1980) in the

provinces of Antioquia and Caldas indicated a shallow continental shelf depositional

environment with local euxinic conditions. Etayo-Serna et al. (2003) described the development

of facies deposited in transitional and shallow middle shelf and external shelf environments

(middle Early Cretaceous and upper Early Cretaceous, respectively) with variations between

sandy and muddy deposits.

Cenozoic Sedimentary Rocks

Honda Group

The Honda Group, which occupies an elongate area that trends north along the Magdelena

River, was defined by Hettner (1892). It is located to the east of the Berlín Project area and

consists mainly of intercalated red sandstones, mudstones and polymict conglomerates

(Barrero and Vesga, 1976). It lies unconformably on Cretaceous sedimentary rocks.

Guerrero (1993) places it in the Middle Miocene interval, a period of development of alluvial

and fluvial continental facies (Cáceres et al., 2003) with predominantly braided and

meandering river deposits.



Mesa Formation

This unit constitutes an upward-fining succession that consists of sandy conglomerates with

clasts of mainly volcanic rocks at its base, overlain by tuffaceous sandstones, lapilli deposits and

volcanic ash. It is discordant over the Honda Group and overlain by recent surficial deposits.

Based on radiometric (Thouret, 1989) and fossil species (Dueñas and Castro, 1981), it is

Pliocene in age.

During the Quaternary, explosive volcanism resulted in volcaniclastic deposits covering a

large part of what is today the Cordillera Central and the Magdalena River Valley. The source

of these deposits is attributed to volcanoes associated with movements on large terrain-

bounding faults, specifically the Palestina Fault in the area of the Berlin Project.

Intrusive Rocks

During the Late Cretaceous period and the early Cenozoic, magmatism in the Cordillera Central

resulted in the emplacement of batholiths and stocks that intrude the pre-Cretaceous and

Cretaceous stratigraphy. The main intrusive bodies are the Antioquia, Sonsón, Ibagué and

Samaná batholiths (Muñoz, 1983; Naranjo, 1983). The Berlin Project lies between converging

faults at the northern end of the Samaná Batholith (Figure 7.1_3). This igneous complex

extends over approximately 150km², measuring about 30km N-S and approximately 8km E-W

on the eastern edge of the Cordillera Central. The complex consists mainly of diorite and

gabbro (~60% of body) with less extensive granodiorites, granites and tonalities (Muñoz, 1983).

Barrero and Vesga (1976) obtained a K/Ar age of 119+/-10Ma on hornblende in the Samaná

Batholith (Barremian – Aptian). Field relationships suggest that the alaskitic component, which

constitutes about 30% of the batholith, was emplaced late in the development of the igneous

complex (Muñoz, 1983). A contact metamorphic aureole extends some 30m to 150m into

enclosing sedimentary rocks adjacent to parts of the batholith that are not fault-bounded.

Igneous rocks of the complex have a homogenous texture with only local structural fabric

development shown by the alignment of crystals such as biotite plates. With the exception of

the alaskite component, the rest of the complex is characterised by an abundance of xenoliths

of gabbro and basalt. Recent drilling has shown that the Cretaceous rocks in the Berlin Project

were intruded by alaskitic dykes and sills and also by granodiorites of Tertiary age.

8.1.2 99B99BStructure

The Cretaceous sequence in which the Berlin Project is located lies at the northern end of the

Samaná Batholith near the point at which two regional faults converge. The San Diego Fault

lies to the east, and the Palestina Fault to the west, of the Berlin syncline. The sedimentary

rocks in the Berlin area are folded into a tight, asymmetric syncline that has a regional vergence

towards the west. The eastern, overturned limb of the syncline is disrupted by bedding – sub-

parallel faults that remove stratigraphy, placing Paleozoic metamorphic rocks on top of black

mudstones which stratigraphically overlie the mineralised zone. These faults have an east-over-

west, thrust-type geometry. The western limb of the Berlin syncline generally dips at a moderate

angle to the east, but is locally disrupted by folding – the geometry of which is east-vergent.



Black mudstones occupy the core of the syncline and have a well developed axial planar

cleavage. The dominant cleavage is east-dipping, axial planar to the principal syncline, but

other cleavage orientations occur in axial planar orientations to minor folds on the west limb of

the Berlin syncline. Cleavage planes are defined by layers of graphite in the most intensely

deformed zones.

8.2 50B50BGeology of the Berlin Project

8.2.1 100B100BLithology & Stratigraphy

Mineralisation at Berlin lies within the Lower Cretaceous Abejorral Formation (Figures 7.2.1_1,

7.2.1_2).

Metamorphic Rocks (εP-Cj)

The metamorphic basement rocks have been correlated with the Cajamarca Complex of

Paleozoic age (Maya and González, 1995). The schists consist mainly of quartz-sericite and

graphitic types. Some quartzites and pelitic sequences are evident in the basement sequence.

Clastic Succession (K-Sj)

The clastic component of the Abejorral Formation fines upward from a discontinuous, basal

conglomeratic unit through clean, well sorted sandstone to black, organic-rich, carbonaceous

mudstones. The conglomerate is reported to average 13m in thickness with a maximum of 30m

(Naranjo, 1983). The conglomerate is matrix- to clast- supported and polylithic and is arranged

in crudely-defined beds, some of which fine upwards into sandstone layers up to 10cm thick.

The conglomerate is poorly sorted with clasts ranging up to 10cm in diameter. The

conglomerate is arranged in a broadly upward-fining manner, with those clasts near the base

being larger and more angular, and those higher in the conglomerate facies being smaller and

of a more rounded nature. The clasts consist of altered basic and intermediate igneous rocks,

metamorphic fragments, milky quartz and medium- to coarse-grained sandstone. In thin section

the matrix of the conglomerate is cemented by clays and carbonates (Naranjo, 1983).

Silicification occurs at the contact with intrusive igneous rocks.

In gradational contact with the basal conglomerate, the overlying sandstone is fine- to medium-

grained and typically some 15-30m thick. Internally the sandstone is arranged in strata of both

tabular and lenticular shape, ranging in thickness from about 10cm to about one metre. Primary

sedimentary structures include cross-bedding, plane bedding and slightly wavy bedding.

Intercalated layers of orange to yellow siliceous siltstone, from centimetres to metres in

thickness, cap some sandstone layers.

Sandstone facies fine upwards through approximately 1m of siltstone into laminated black

mudstone. Recent drilling shows that the mudstone exceeds 300m in thickness. The mudstone

shows parallel lamination and some parts of the sequence is interstratified with grey siliceous

mudstone approaching chert in texture in layers from one to ten centimetres thick. Bivalve

fossils occur in the lower part of the sequence and ammonite fossils are sparsely distributed

throughout the upper part. Pyrite occurs as fine disseminations which usually follow the primary

sedimentary laminations in the mudstone, as well as in nodules. The mudstone is bituminous in

part and contains graphite in others.



Figure 7.2.1_1 Geology of the Berlin Project



Figure 7.2.1_2 Generalised Stratigraphic Column of the Berlin Project



Carbonate Facies (K-Be)

Drillhole intersections suggest that the upper part of the sandstone facies transitions laterally into

carbonate facies beneath the mudstone sequence over most of the southern part of the Berlin

syncline. The carbonate unit is tabular and averages 15m thick in the southern 3km of the Berlin

syncline. The carbonate is made up of a remarkably persistent facies sequence that has been

consistently identified in drillholes drilled throughout the southernmost 3km of the Berlin syncline.

The lowermost carbonate facies is a unit that clearly has an erosional contact into the underlying

fine-grained sandstone in some drillhole intersections. The lowermost carbonate unit consists of

a matrix-supported fragmental unit. Clasts are up to 60cm in diameter and many contain soft-

sediment deformation features in which flame-like structures are evident with embayments of

matrix into the clasts and vice versa. Clasts vary in composition and include shelly, fine-grained

carbonate facies, carbonate mudstones and coarse granular facies (Figure 7.2.1_3). The matrix

varies from massive mud to granular material and shells and shell fragments are suspended in

the matrix. This fragmental unit is variable in thickness and is typically one to three metres thick.

This matrix-supported fragmental rock fines upwards with the upper part of the unit exhibiting

fine parallel lamination.

Figure 7.2.1_3

Coarse Fragmental Facies in Drill Core

A – Bivalve shell, B – Gastropod shell, C – Soft-sediment deformation features – flame-like structures,

D – Fine-grained mud clasts, E – Coarse-grained clasts, F – Crinoid stem in fine-grained mud-dominated matrix

The overlying unit consists of stacked, upward-fining sequences that lie on scoured surfaces.

These upward-fining units typically commence with a granular unit that contains equant shell

fragments (Figure 7.2.1_4). These granular units are typically massive in texture and fine

upward into sandy carbonate facies that contain abundant shell fragments which are exclusively

from bivalves. The bivalve fragments are typically small in the lower part of the sandy carbonate

facies and generally become larger as the matrix fines upwards into carbonate mud. The fine

sandy carbonate facies tends to be plane laminated, culminating in a conspicuously laminated

carbonate mudstone (Figure 7.2.1_4) which constitutes an easily correlated marker unit. Some

intersections of the finest facies show wavy lamination.



Figure 7.2.1_4 Drillhole DDB-013 Sedimentary Sequence



These stacked, upward-fining layers constitute a unit that varies between 1m and 10m thick,

and grades upwards into black mudstone. Some intersections show rare fine sandy

carbonate laminae that fine upwards within the mudstone sequence.

The fossil assemblage dates the sequence at Valanginian – Hauterivian (Muñoz, 1983).

Interpretation of Environment of Deposition of the Sedimentary Cretaceous Sequence

The environment of deposition of the Cretaceous sequence in the Berlin Project area can be

interpreted in terms of gravel-dominated alluvial fans arranged along a scarp, with volcanic and

metamorphic material shedding from the paleo-high into a basin that is locally developed

adjacent to the Palestina Fault. Erosion of the scarp with time leads to lower-energy fluvial

environments and the development of sand facies overlying the polylithic conglomerates.

Continued subsidence results in a transition to shallow marine sands that contain gastropods

and worm burrows. In the southern part of the Berlin syncline, these sands transition laterally

and vertically through silt and subsequently mud facies that contain marine fauna. In the central

part of the Berlin syncline, sandstone facies transition laterally and vertically through siltstones to

carbonate facies, the basal part of which has the characteristic of a debris flow deposited in

channel-like features scoured into the underlying marine clastic facies. The boulder-sized

carbonate fragments are transported in a matrix of mud, which fines upwards into a laminated

carbonate mudstone facies – typical of debris flows in subaqueous environments. The overlying

sediments are carbonate sands and muds that contain fossil fragments that consist almost

exclusively of bivalve shells that increase in size upwards through the sequence. There is

evidence of erosional contacts overlain by a repeat of similar facies higher in the sequence. The

carbonate facies are similar in structure to Bouma sequences in turbidites (Bouma, 1962) in

which upward-fining facies are arranged in a predictable order on erosional scour surfaces

formed in response to mass flows slumping into deep, tranquil waters.

Volcaniclastic Deposits (Qvlc)

Volcaniclastic deposits are distributed throughout the project area and were erupted from the

caldera at San Diego located in the northeastern part of the Berlin Project area. Flow

structures are observed in sectors close to Lake San Diego. Lithologically, these ejecta are

classified as tuffs made up of fragments of metamorphic and sedimentary rocks from the

underlying strata in a matrix of volcanic ash and lapilli.

Fluvial deposits (Qfa)

The most important accumulations of fluvial deposits are associated with the Manso river and

the Santa Marta creek. Alluvial terraces are made up of polymict, clast-supported

conglomerates with rounded to sub-rounded clasts of igneous, metamorphic and sedimentary

rock. The matrix varies from very coarse-grained to very fine-grained sand.



Intrusive Rocks (K2N-Sm)

Igneous rocks occur as plutons, sills and dykes, which intrude both the metamorphic

basement and the Cretaceous sedimentary units.

The alaskite that intrudes the southwestern part of the Berlin syncline has an equigranular

texture and is composed of plagioclase, quartz and minor biotite. Two granodiorite stocks

outcrop on the eastern margin of the Berlin syncline. The granodiorite consists of coarse

subhedral crystals of plagioclase, quartz and amphibole. These stocks are thought to

represent a different magmatic pulse from the alaskite in the west.

Andesitic and dacitic dykes and sills encountered in the area are locally porphyritic, with

subhedral plagioclase phenocrysts in a felsic matrix.

8.2.2 101B101BStructure

Regional mapping shows that Cretaceous rocks in the Berlin area are folded into a tight, keel-

shaped synform whose eastern limb is steep to over-turned, and dips to the east, while the

western limb is generally moderately inclined to the east. This west-verging synform is

characterised by an axial planar cleavage which is strongly developed in the black mudstone

sequence.

Project-scale mapping and drilling shows that Palaeozoic schists are in contact with the black

mudstone unit in some sectors of the eastern limb of the Berlin syncline. Trenches in these

areas show west-verging thrust faults, which are consistent with structures that eliminate

stratigraphy that have been intersected in recent drilling. These data indicate that the

Palaeozoic metamorphic rocks have been thrust over the Cretaceous sediments, eliminating

the clastic and carbonate sequences in these areas. The structure of the Berlin syncline, as

interpreted from recent mapping, trenching and drilling, is shown in a sequence of cross

sections in Figures 7.2.2_1 and 7.2.2_2.

Trenching in the southwestern part of the Berlin syncline reveals important features of the

west limb in that area. Trench TB28 is located on the western margin of the Section 1

(Figure 7.2.2_1) and exposes several ramp-flat features that separate relatively flat-lying

footwall strata with more steeply inclined hanging wall beds. The ramps detach onto a fault

that lies at the upper contact of a tan-coloured, clay-rich unit that is located in the footwall of

the mineralised zone in most trenches. These features indicate that the mineralised zone in

TB28 is located within a west-verging duplex structure on the west limb of the Berlin syncline

(Figure 7.2.2_3).

A section exposed immediately southwest of drill platform P1 (Figure 7.2.2_1) shows that the

sandstone beneath trench TB28 is overturned, forming the west-dipping limb of a fold that has

an easterly vergence, opposite to that of the regional syncline. These data are consistent with

the depth of the intersections made in drillholes BDD001-to BDD003 as shown in cross

Section 1 of Figure 7.2.2_1.



Figure 7.2.2_1 West-East Cross Sections (1 to 4) Through Berlin Syncline and Plan View



Figure 7.2.2_2 West-East Cross Sections (5 to 12) Through Berlin Syncline

Refer to Figure 7.2.2_1 for plan view location of cross sections



Figure 7.2.2_3 Geological Profile of Trench TB28 Located on the West Limb of the Berlin Syncline on Section P1

West to east cross section. The fine-grained sandstone is the principal mineralised unit

8.3 51B51BMineralisation

8.3.1 102B102BMethod of Study

The description of the host rocks and mineralisation at Berlin relies on work that was carried

out as described below, and adds to the original petrographic work by Castaño (1979),

summarised in the previous NI 43-101 report on the Berlin area (Spencer & Cleath, 2010).

U3O8 Corp. commissioned petrographic and microprobe studies with a view to investigating

mineral textures and to characterise the uranium-bearing, and associated, phases.

Jim Renaud of Renaud Geological Consulting used transmitted and reflected light microscopy

and a microprobe to study and analyse samples from trenches and drillholes in the southern

part of the Berlin syncline. A composite sample taken from mineralised drillhole intersections

was investigated by XRF, XRD and scanning electron microscopy (“SEM”) and backscattered

electron microscopy (“BSE”) by ANSTO Minerals (ANSTO, 2011). SEM and reflected and

transmitted light microscopy form part of a postgraduate study (Caceres, in prep.). The

results of these studies are summarised below.

8.3.2 103B103BDescription of the Mineralised Zone

Sandstone

Mineralisation at Berlin is stratiform and is located beneath a carbonaceous mudstone unit.

Trenching has traced the mineralised zone over a distance of 8.5km on the east limb of the

syncline and 3.5km on its west limb.



The mineralised zone exposed in trenches typically has a tan-coloured, clay-rich footwall that

varies from 40cm thick to several metres thick and contains several hundred ppm U3O8

(Figure 7.2.2_3). The tan-coloured layer is typically overlain by a mature, well-sorted fine- to

medium-grained calcareous sandstone that contains variable amounts of interstitial

bituminous hydrocarbon. This mature sandstone is consistently mineralised. The clean, well-

sorted sandstone fines upward over about 30cm into a laminated mudstone-siltstone unit in

which uranium grades typically drop to several tens of ppm U3O8 over about a metre from the

upper contact of the sandstone. The thickness of the mineralised zone varies between 1m

and 6.7m and averages approximately 2.3m thick.

Fracture-fillings in the sandstone near the southern closure of the synform have typically

green to pale blue to orange, botryoidal coatings that have been identified as variscite

(AlPO4.2H2O) and childrenite ((Fe,Mn)AlPO4(OH)2.H2O) (Toronto Museum, 2010). These are

typically developed in the weathering environment from the primary phosphate minerals, and

occur within interstices in the sandstone. Similar fracture-fill secondary mineralisation extends

some 30m into the sandstones that underlie the mature sandstone that hosts the stratiform

mineralisation. The average grade of the footwall fracture zone in a profile made from trench

TB28, located approximately 100m west of platform P1 (Figure 7.2.2_1), is ~400ppm U3O8. It

is not clear from the available field data whether the footwall fracture zone in the vicinity of

TB28 is restricted to that geographic area or whether it is related to dilation in the axial zone of

the syncline.

A similar mineralised profile is observed in core from drillholes that cut the sandstone – hosted

mineralisation.

Carbonate Rocks

The principal uranium-mineralised zone in the carbonate host-rocks is tightly constrained to a

specific facies that was intersected in most of the drillholes drilled in the southern part of the

Berlin syncline. This principal unit is typically sandwiched between two lower-grade units,

from which grades decrease sharply into the foot wall and hanging wall.

8.3.3 104B104BComposition and Textures of the Host Rocks

Weathered Sandstone

Petrographic studies show that the sandstone samples taken from trenches are composed

mainly of coarse quartz grains with minor magnetite and hematite, barite and minor chromite

(Renaud, 2010a).

Interstices are filled by apatite, Fe-Al-Ti-Cu-Ca-Cr-bearing phosphates, roscoelite (a V-Ba mica),

Y-phosphates including churchite, monazite (REE-phosphate) and numerous U-bearing

phosphates of the autunite (Ca(UO2)2(PO4)2.10-12H2O) and meta-autunite subgroups (Renaud,

2010a). Mineralisation was observed to be parallel to bedding planes and planes of weakness

in the rock.



Some sandstones consist of micro - mosaic-brecciated quartz grains. These are breccias,

that if the matrix was removed, the fragments would fit together like a jig-saw puzzle to

reconstitute the original grains (Figure 7.3.3_1). The matrix between these fragments

consists of fine-grained apatite, churchite, Fe-Al-phosphate +/- U, fine-grained zircon, Fe-Al-

phosphate, roscoelite and Fe-Al+/-Ca-Ti-V-Cr-phosphate.

Figure 7.3.3_1

Sandstone

A common texture in these sandstones is mosaic-brecciation of the larger grains (the fragments fit together like a jigsaw

puzzle) that are cemented with yellow cacoxenite and black-green V-Ba mica (Roscoelite).



In one weathered sandstone sample, the rock is dominated by quartz grains with interstitial

matrix consisting of apatite, Ba-V-mica, and phosphates defined by alternating bands of black

and yellow. The black bands are dominated by quartz with an interstitial matrix of apatite, Ba-V-

mica, and minor Fe-Al-phosphates (cacoxenite). The interstitial apatites commonly contain fine

grained REE-phosphate (monazite). The yellow bands are dominated by quartz with interstitial

matrix dominated by apatite, cacoxenite and lesser V-bearing mica (Figure 7.3.3_2).

Figure 7.3.3_2

Sandstone Photomicrographs In Plane and Crossed Polarised Light

Photomicrographs in plane and crossed polarised light illustrating a complex domain of yellow cacoxenite

intergrown with pale green Al-Ba-Fe-Si-Ca-U-phosphate. From Renaud, 2010a.

Elongate quartz grains with yellow cacoxenite define the pervasive foliation in some samples.

Cacoxenite in another sample of sandstone occurs in association with intergrowths of Fe-Al-

V-phosphate and U-V-Al-Fe-Ca-phosphates (Renaud, 2010a).



An early generation of quartz veinlets is folded with primary layering (bedding) in one sample.

These veinlets are cross-cut by a later generation of stockwork veinlets which show small-

scale dislocations, but are not folded (Renaud, 2010a). This suggests that there were two

phases of silicification of the sandstone – the early one being pre-folding, and the more

pervasive one being post-folding.

Fresh Sandstone

In samples of the sandstone facies from drillholes DDB1 and DDB3, there are coarse clastic

domains dominated by quartz with interstitial domains of mica-apatite-chlorite-pyrite-rutile-

zircon in contact with more apatite-rich domains which seal finer-grained quartz, mica, iron-

oxide, zircon, rutile, and a host of metals (Renaud, 2010c). These apatite-rich domains host

such metals as silver-poor tetrahedrite, Ni-S (millerite), pyrite, sphalerite, and fine-grained U-

Ti-bearing minerals. The tetrahedrite in a sample from 110.85m depth in drillhole DDB1 is a

unique mineral in the petrographic studies undertaken to date because there is no silver peak

in the EDS spectra.

Pyrophyllite was observed to be replacing mica and quartz grains in drillhole DDB3, where it

constitutes up to 20% of the interstitial mineralogy between quartz grains. The pyrophyllite is

occasionally associated with pyrite grains, but has not been noted with any other metals in

these sections.

Carbonate Rocks

The results of analysis by XRD of the composite sample obtained by blending carbonate

facies material from 15 drillholes (Section 13.3,) are shown in Table 7.3.3_1, below (ANSTO,

2011). Major mineralogical phases present were confirmed by X-ray diffraction with

quantification undertaken using Siroquant. Major and minor mineralogical constituents were

also assessed using a scanning electron microscope (SEM) equipped with an energy

dispersive system (EDS). The composite Berlin sample is dominated by calcite with lesser

amounts of fluorapatite and quartz (Table 7.3.3_1). Minor phases detected by XRD include

muscovite, dolomite, pyrite, chlorite and sphalerite.

Table 7.3.3_1

Berlin Project

Relative Concentrations of Minerals in the ANSTO1 Composite Sample of Carbonate Facies

Mineral Chemical Formulae Wt (%)

Calcite CaCO3 57.7 Fluorapatite Ca5(PO4)3F 18.2 Quartz SiO2 15.8 Muscovite (K0.82Na0.18)(Fe0.03Al1.97)(AlSi3)O10(OH)3 2.6 Dolomite CaMg(CO3)2 2.5 Pyrite FeS2 1.9 Chlorite (Mg,Fe)5Al(Si3Al)O10(OH)8 0.9 Sphalerite ZnS 0.4



Bedding in fine-grained carbonate facies from drillhole DDB7 is defined by calcite, quartz and

pyrite (Renaud, 2010b). Circular fossil fragments (crinoid stems?) are defined by circular

domains of calcite and transparent apatite and some are enclosed by magnetite. Framboidal

pyrite occurs in the carbonate rock.

Two populations of apatite were observed in clastic domains within the carbonate facies

(Table 7.3.3_2, Figure 7.3.3_3). Domains of quartz grains and clear apatite contain fine-

grained pyrite inclusions that do not extend beyond the domain boundaries suggesting that an

early phase of pyrite accompanied quartz-apatite growth in clastic layers adjacent to

limestone facies. In addition, the clear apatite grains are pitted and their embayed surfaces

and truncated pyrite inclusions are overgrown with the later assemblage of black apatite,

calcite, pyrite and sphalerite (Figure 7.3.3_4).

Figure 7.3.3_3

Carbonate Facies Photomicrographs from a Clastic Layer

Photomicrographs from a clastic layer within the carbonate facies illustrating the black, isotropic apatite intergrown with

interstitial quartz and calcite.

From Renaud, (2010b).



Figure 7.3.3_4 Backscatter Images of Fractures

The fractures are filled with black Mg-chlorite, black isotropic apatite, quartz, pyrite and sphalerite. From

Renaud, (2010b).

Table 7.3.3_2

Berlin Project

Comparison of the Composition of the Two Types of Apatite in Carbonate Facies Rocks in Drillhole DDB7

Apatite Species CaO (%) P2O5 (%) F (%) CO2

Clear apatite 58.19 42.49 0 0 Black isotropic fluorapatite 52.06 39.04 2.77 Present



8.3.4 105B105BNature of the Mineralisation

Uranium

Uranium is present in the three major uranium phases including:

Uraninite/pitchblende (UO2). Uraninite/pitchblende is the most abundant of the uranium

phases in the southern part of the Berlin Project, occurring as disseminated, very fine

particles that typically range between 5µm and 10µm in size, with some rare grains

reaching 30µm in diameter (ANSTO, 2011). Uraninite is closely associated with graphite,

forming in small, round masses at the margins of graphite flakes and also on cleavage

planes within the graphite (Figure 7.3.4_1; Caceres, in prep.). Uraninite also occurs at

grain boundaries of calcite, apatite and quartz (Figure 7.3.4_2) and within grains of Stage 2

apatite and in calcite (Figure 7.3.4_3).

Coffinite (U(SiO4)1-x (OH)4x). Coffinite most often occurs as small (<10μm) particles hosted

by mica. The micaceous hosts are often located in the interstices between calcite, apatite,

quartz and sulphide grains.

Brannerite ((U,Ca,Ce)(Ti,Fe)2O6). Brannerite is the least abundant of the uranium phases

present. It tends to occur as multi-crystalline laths.

Figure 7.3.4_1

Backscatter Images of Uranium Bearing Minerals

Uranium-bearing minerals are the bright spots on the contacts of graphite grains and along fractures through the graphite

(Caceres, in prep.). U: uranium inclusion; Gr: Graphite.

Most of the uraninite/pitchblende grains exhibit some alteration; suggested by their apparently

multi-crystalline cores and bright rims. One rare example was found of uraninite/pitchblende

veining in a xenotime-like phase.



Figure 7.3.4_2 BSE Image of the Distribution of Uraninite Particles at Grain Boundaries within the Carbonate Facies

Source (ANSTO, 2011).

Figure 7.3.4_3

BSE Image Showing Pale Uranium Mineral Inclusions In Stage 2 Fluorapatite

(U = uraninite; Ap (St 2) = Apatite Stage 2; Gr = Graphite; M3 = Matrix)

Source (ANSTO, 2011).



Graphite

In the drillhole core, graphite is observed to form thin, often striated layers in the plane of the

cleavage. Microscopic investigation shows the partial replacement of organic matter (bitumen)

by graphite (Figure 7.3.4_4). Graphite forms elongate masses that are either amorphous or

contain internal cleavage planes similar in orientation to those in the surrounding host rock

(Figure 7.3.4_4a). Graphite also occurs in the matrix of microbreccias (Figure 7.3.4_4b).

Figure 7.3.4_4

Graphite In Reflected Light Microscopy Images

Graphite occurs as amorphous masses in the plane of the cleavage, a) and also is matrix fill in

microbreccias; b) Sphalerite follows the same oriented patterns as graphite. (Gr =Graphite; ZnS = Sphalerite; M.O = Bitumen).

Apatite

Samples from carbonate facies contain two populations of apatite (Renaud, 2010b):

Type 1: this is clear apatite that has a low fluorine content and contains inclusions of

pyrite. Colourless apatites have a distinct composition relative to that of the dark apatites

described below (Table 7.3.3_2).

Type 2: this black, isotropic fluorapatite occurs in fractures and interstitial to quartz

grains. Cross-cutting relationships show that this apatite formed after the clear apatite.

Black apatite is commonly associated with calcite and Mg-chlorite with pyrite, sphalerite

and millerite (Ni-S), and contains fine-grained inclusions of a uraninite. Isotropic apatite

has a high fluorine content and a carbonate component. Cryptocrystalline Type 2 apatite

or collophane coats embayed quartz grains and grades outwards into crystalline apatite

(Figure 7.3.4_5).

Calcite

Microscopic investigations reveal various phases of calcite in the Berlin host rocks:

Crystalline calcite and apatite are observed replacing fossil fragments.

A phase of calcite occurs in cross-cutting relationships with earlier calcite. Cross-cutting

calcite occurs with Type 2 apatite, magnesian chlorite, pyrite, sphalerite, millerite and

uraninite.

A phase of calcite also partially replaces Type 2 apatite (Figure 7.3.4_6) and graphite

(Figure 7.3.4_7).



Figure 7.3.4_5 Evidence of Stage 2 Phosphate Precipitation in Transmitted Light Microscopy Image

Transmitted light microscopy image showing embayed quartz grain partially overgrown with collophane, grading outward,

away from the quartz grain, into crystalline apatite (Cal = Calcite; Ap (St2) = Apatite Stage 2).

Figure 7.3.4_6

Backscatter Image of Branching Stringers of Calcite

Replacement of Type 2 apatite by calcite (Apt = Apatite (Type 2) and Cal = Calcite)



Figure 7.3.4_7 Backscatter Image of Replacement of Graphite by Calcite

Note the uraninite grains located on fractures through the graphite.

(U = uranium-bearing mineral, BT = graphite and Cal = calcite)

Sulphides

Sulphides are present in a various forms in the mineralised zone at Berlin. Pyrite and

sphalerite are the most abundant of the sulphide minerals.

One phase of pyrite occurs as fine framboids (Figure 7.3.4_8). Another phase of pyrite occurs

with and within apatite. Pyrite is observed in fractures in calcite with chlorite, black isotropic

apatite, quartz and sphalerite. Calcite-quartz-apatite domains in carbonate facies rocks are

cemented by a later phase of calcite with zircon, monazite and a later phase of pyrite.

Sphalerite is commonly spatially associated with graphite and, in some instances, lies within

cleavage planes in graphite. Anhedral sphalerite also occurs clustered with other sulphides

associated with quartz (Figure 7.3.4_9). Ni-As sulphides are observed to partially replace

sphalerite.

Ni-sulphides (millerite) and Fe-Ni-sulphides (pentlandite) are present in less abundance than

the sphalerite and pyrite (Figure 7.3.4_9).

Rare Earth Minerals

REE-fluoro-carbonate (bastnaesite), containing La, Ce and Nd, is present as a rare phase in

quartz-mica-sulphide – bearing parts of the carbonate facies.

Monazite, containing Ce, La and Nd, also occurs as a rare phase. The monazite is typically

present as sub to anhedral particles hosted by quartz and calcite.



Figure 7.3.4_8 BSE Image Showing Fe-Ni Sulphide (Pentlandite) and Framboidal Pyrite in the Carbonate Facies

Figure 7.3.4_9

BSE Image Showing Ni-As Sulphide (Gersdorffite?) in Contact with Quartz and Calcite from the Carbonate Facies



8.3.5 106B106BAlaskite

The alaskite has a conspicuous granophyric texture (Figure 7.3.5_1) and contains numerous

minerals commonly seen in the adjacent sandstones (Renaud, 2010a).

Figure 7.3.5_1

Photomicrographs of Coarse Plagioclase

Photomicrographs of coarse plagioclase, some with well preserved albite twinning and relict growth zones. The albite has been

pervasively altered to K-feldspar.

The alaskite contains plagioclase with well-preserved albite twinning and relict concentric

growth zones intergrown with quartz and relict igneous mica that has been altered to sericite.

The plagioclase grains have been pervasively altered to Ba-K-feldspar. Areas that are

interstitial to the plagioclase-quartz grains contain finer-grained quartz and K-feldspar with

strong granophyric textures. The altered albite commonly contains inclusions of zircon and

rutile and the K-feldspar hosts zircon and Y-phosphate inclusions. Zircon also occurs with the

quartz grains. REE, Y-phosphate and monazite are common throughout the thin section

studied and are associated with rutile, K-feldspar and albite.



8.3.6 107B107BSummary of Mineral Textures & Observations Regarding Paragenesis

Sedimentary facies and paleontology indicate that the host rocks of the mineralisation at

Berlin were deposited in a shelf-type, pelagic environment in the early Cretaceous. The

occurrence of framboidal pyrite in parts of the carbonate sequence and in the overlying black

mudstone is characteristic of biogenic activity and hence this phase of pyrite is likely to have

precipitated at the time of deposition or at least prior to lithification.

Domains of quartz grains, clear apatite and fine pyrite within the carbonate facies are indicative

of an early phase of pyrite precipitation with Type 1 apatite in more clastic components of the

rock (Renaud, 2010b). Type 1 apatite and calcite partially replace shell fragments.

Embayed and pitted quartz grains tend to be elongate in the plane of the cleavage suggesting

that dissolution of silica occurred during cleavage formation. These embayments in quartz are

encrusted by crystalline calcite and cryptocrystalline collophane, which grades outward, away

from the quartz grain, into crystalline, Type 2 apatite (Caceres, in prep.). Silica consumption in

some sandstone facies is marked by the partial replacement of quartz by pyrophyllite (Renaud,

2010a).

Cleavage planes are defined by graphite, mica and to a lesser extent, by cacoxenite

intergrown with roscoelite, apatite, zircon and uranium-bearing phosphates. Cacoxenite

extends from breccia filling into the cleavage. Uraninite occurs on the margins of graphite

flakes and on cleavage planes within the graphite. Uraninite also occurs in Type 2 apatite and

in some calcite crystals.

Parts of the carbonate facies composed of calcite, quartz and fine-grained apatite (Type 2)

are cemented by a later phase of calcite with zircon, thorium-bearing monazite and pyrite

(Renaud, 2010b).

Calcite overgrows graphite cleavage and replaces Stage 2 apatite (Caceres, in prep.),

suggesting that there is a post-tectonic calcite pulse.

Uraninite grains cluster on the margins of graphite that lies within the plane of the axial planar

cleavage. The occurrence of graphite is significant in terms of defining the timing of uranium

mineralisation since the graphite appears to have formed from bitumen – the black mudstones

are known source rocks for hydrocarbons in Colombia, Ecuador and Peru. If this was the

case, in order to have formed bitumen, the source rocks would have passed through the oil

window, to the extent that oil transformed to bitumen and subsequently to graphite. Graphite

is one of the minerals that defines the axial planar cleavage in the Berlin synform, which

suggests that the graphite formed during folding of the sequence, which is likely to have been

during basin inversion which commenced in the Paleogene and was most intense in the

Miocene and Pliocene (Cediel et al., 2003). For uraninite to lie along the margins of graphite

grains, as well as in fractures within graphite, suggests that at least this phase of uranium

mineralisation occurred contemporaneously with, or after, folding that formed the Berlin

syncline.



Quartz grains are elongated in the orientation of the cleavage – and the lack of strain

shadows in the quartz grains, combined with etched margins, shows that the elongation of the

quartz grains was principally by dissolution. Minerals occupying embayments into quartz

grains, occupying sites of dissolution, are calcite, Stage 2 apatite, chlorite and sulphides.

These textural relationships suggest that these minerals also precipitated during or after

cleavage formation.

Some quartz grains in the sandstones have fracture patterns which do not extend beyond

grain margins, providing evidence of an early fracture event. These same grains have

fractures that do extend out into the matrix and hence were formed later than the fractures

which do not extend beyond grain boundaries. Only the later fractures are filled by yellow

cacoxenite, black-dark green roscoelite, apatite, zircon, and U-phosphates, indicating that

these minerals precipitated not in the first fracture event, but in the second.

It is noteworthy that the alaskitic rocks that lie immediately to the west of the Berlin syncline

contain zircon and yttrium phosphates that have similar textures to those minerals found in

mineralised parts of the adjacent sedimentary sequence.



9 7B7BDEPOSIT TYPES

9.1 52B52BHistoric Perspective

Minatome’s exploration work led to the conclusion that the mineralisation in the Berlin Project

is syngenetic – an accumulation of uranium in a phosphatic shale unit similar to the Alum

Shale in Sweden that contains uranium with vanadium and various other elements (Mining

Journal, 2009).

9.2 53B53BAnalogous Deposits

The closest analogy to the Berlin Project is the Santa Quiteria phosphate-uranium deposit in

northeastern Brazil. The Santa Quiteria Batholith is a Neoproterozoic continental magmatic

arc composed of several granitoid types (Fetter et al., 2003). This deposit contains

approximately 200mlbs of uranium and the associated phosphate occurs in both stratiform

bodies and in cross-cutting veins in a limestone host rock near alkaline intrusives. The

reported non-CIM, non- NI 43-101 compliant resource is 80 million tonnes at 11% phosphate

and 0.1% U3O8. No genetic relationship has yet been demonstrated between the intrusive

and mineralisation at Santa Quiteria.



10 8B8BEXPLORATION

10.1 54B54BResponsibility for Exploration

All of the recent exploration described in this report has been done by employees and

consultants of Gaia Energy (Colombia) Ltd, an indirectly held, wholly owned, subsidiary of

U3O8 Corp. Only U3O8 Corp’s assay data and drill results have been used for the resource

estimate.

10.2 55B55BApproach

Due to the stratiform nature of the mineralisation at Berlin, the principal objective was to define

the extent and consistency of the known mineralised layer through trenching and drilling.

10.3 56B56BTrenching

Trenches were excavated by hand on outcropping mineralisation on the flanks of the Berlin

Syncline. Trench sites were identified in two principal ways:

With the use of geological maps made by Minatome in the late 1970’s and early 1980’s

that indicated areas of outcropping mineralisation, coupled with the help of field

assistants whom had worked for Minatome in its prior exploration of the area; and

Reconnaissance transects made roughly perpendicular to the axis of the syncline in which

outcrops and subcrops of mineralisation were identified with hand-held spectrometers.

Once the anomalously radioactive stratum was identified, the trench was cut perpendicular to

strike. Sample locations were then defined based on lithology and levels of radioactivity.

Radioactivity was measured with a hand-held GR 135 spectrometer.

To date, 38 trenches have been excavated. The majority of the trenches are located on the

more accessible southern part and eastern flank of the syncline, where mineralisation has

been shown to occur over a strike distance of 8.5km. A summary of assay results obtained

from the trenches is shown in Table 9.3_1 and Figures 9.3_1 and 9.3_2 display the location of

the trenches.

10.4 57B57BConclusion

The trench information has supplied further information as to both the nature of mineralisation

in the project area, and to the quantum of the uranium mineralisation as indicated by the

historical work. The trench assay results further support the continued drilling of the deposit.



Table 9.3_1

Berlin Project

Summary of Trench Assay Results

Trench ID Estimated True

Thickness U3O8 (%)

U3O8 (lb/st)

V2O5 (%)

P2O5 (%)

Mo (ppm)

Re (ppm)

Ag (ppm)

Ni (ppm)

Zn (ppm)

Y2O3 (ppm)

Nd2O3 (ppm)

TB-003 1.35 0.085 1.88 0.862 6.1 173 0.1 0.9 25 114 264 429

TB-001 1.34 0.114 2.51 1.066 3.9 769 0.0 1.7 13 284 135 216

TB-023 1.90 0.069 1.53 0.908 12.2 60 0.0 0.9 430 612 235 933

TB-002 2.32 0.172 3.79 0.887 4.9 140 0.0 1.6 37 40 205 853

TB-019 2.24 0.066 1.46 1.121 3.0 165 0.0 1.4 196 480 265 1341

TB-010 0.88 0.065 1.43 0.984 15.7 126 0.0 0.4 409 511 171 860

TB-011 0.99 0.039 0.85 0.821 3.6 16 0.0 0.7 57 115 353 1145

TB-009 2.64 0.134 2.95 0.828 14.6 146 0.0 0.4 43 85 237 840

TB-026 0.42 0.068 1.50 0.956 12.3 343 8.0 3.9 1873 2908 71 336

TB-025 0.88 0.081 1.79 1.011 13.0 216 0.0 0.5 13 6 249 925

TB-020 1.32 0.145 3.19 1.323 16.6 27 0.1 2.9 24 33 215 921

TB-000 0.73 0.114 2.51 0.995 2.3 142 0.1 2.3 432 712 245 1078

TB-004du 2.55 0.093 2.05 1.100 12.8 88 0.0 0.7 54 116 257 851

TB-027 5.47 0.110 2.41 0.855 17.4 238 0.0 1.9 783 2213 223 971

TB-004 1.68 0.083 1.83 1.075 3.7 172 0.0 0.3 278 570 241 1180

TB-016 1.53 0.321 7.06 1.382 25.9 1584 35.9 10.6 7063 3109 283 1252

TB-029 0.90 0.197 4.33 0.830 8.7 120 0.1 9.3 133 17 242 752

TB-028 2.72 0.134 2.94 0.464 6.7 153 0.0 1.2 33 138 163 683

TB-013 0.50 0.077 1.69 0.963 19.6 81 0.1 0.7 382 502 249 960

TB-006 1.86 0.100 2.21 0.701 13.0 40 0.1 1.3 19 39 197 621

TB-014 0.71 0.063 1.38 0.591 4.9 58 0.2 1.7 45 90 26 123

TB-012 2.88 0.081 1.79 0.406 3.7 216 0.0 3.6 42 80 127 454

TB-031 0.59 0.048 1.05 0.295 7.1 36 0.1 1.0 27 53 150 505

TB-031 0.92 0.050 1.09 0.632 10.1 102 0.1 1.1 22 29 125 448

TB-032 0.80 0.063 1.38 0.630 8.0 74 0.0 1.3 312 191 155 660

TB-032 0.59 0.050 1.09 0.420 2.3 38 0.0 10.0 324 129 21 133

TB-034 1.90 0.100 2.20 0.563 5.7 99 0.1 3.4 77 322 73 333

TB-036 0.70 0.035 0.77 0.630 0.9 141 0.0 9.8 362 415 0 0

TB-008 1.70 0.075 1.65 0.745 11.4 30 0.0 0.8 17 34 162 733

TB-005 2.01 0.136 2.99 0.568 12.2 105 0.0 0.9 32 74 176 637

TB-021 2.20 0.142 3.13 0.862 15.1 34 0.1 1.9 209 262 264 1016

TB-033 2.47 0.129 2.84 0.742 18.1 53 0.1 1.1 11 18 321 1110

TB-022 2.39 0.085 1.87 0.935 8.7 179 1.8 2.8 30 48 107 332

TB-018 1.95 0.064 1.41 0.684 12.0 38 0.0 0.4 50 73 249 755

TB-037 2.62 0.114 2.51 1.055 9.5 292 0.0 0.8 537 481 424 1871

TB-038 2.08 0.086 1.90 0.932 13.3 70 0.0 0.6 79 243 176 729

Summary assay results from the mineralised intervals exposed in trenches in the southern part of the Berlin syncline with 0.4% U3O8 cutoff grade.



Figure 9.3_1 Geological Map of Southern Berlin Syncline Showing Location of Trenches Completed in 2010 and 2011



Figure 9.3_2 Geological Map of Northern Berlin Syncline Showing Location of Trenches Completed in 2010 and 2011



11 9B9BDRILLING

11.1 58B58BDrill Programs

Kluane Drilling Ltd. (‘Kluane’) of Whitehorse, Canada commenced drilling at Berlin on

15th October 2010, and a second rig commenced on 10th April, 2011. Both rigs left site on

26th October, 2011. The monthly drilling rates are summarised in Figure 10.1_1. Kluane used

man-portable KDHT-1000, wireline rigs that the drill company has designed and manufactured

in-house. The core drilled is B-thinwall (BTW) that has a core diameter of 42mm, comparable

with the 47.6mm diameter of NQ core, and N-thinwall (NTW) that has a core diameter of

57mm, comparable with HQ at a 63.5mm diameter.

Figure 10.1_1

Monthly Drilling Rate In Berlin

Downhole radiometric analysis was done with a Mount Sopris probe manufactured by Mount

Sopris Instruments and calibrated at that company's Grand Junction, Colorado facilities. On

completion of each drillhole, the probe was lowered to the bottom of the hole on a cable and

the radioactivity was measured at 10cm intervals as the probe was winched up the hole. Data

from the probe was downloaded at the field camp, analysed and stored in a database.

The objective of the initial drilling of approximately 1,500m was to test the depth extent of

mineralisation beneath the mineralised trenches and to define the shape of the syncline within

which the mineralisation occurs. A crucial component of the drilling was also to obtain fresh

samples with which to initiate metallurgical testwork. Environmental impact was minimised by

drilling multiple holes, on section, from one drill pad.



Given the success of the initial drilling in intersecting consistent mineralisation, the program was

extended, culminating in the completion of 82 drillholes for 18,523m. The location of the

platforms from which the holes were drilled is shown in Figure 10.1_2 and the list of holes drilled

from each platform is shown in Table 10.1_1. A summary of assay data from the drill intercepts

is shown in Table 10.1_2.

Figure 10.1_2

Locations of Drill Platforms

Drill Platforms are labelled “P” from which 82 diamond drillholes were completed.

Trench location are shown as black rectangles



Table 10.1_1

Berlin Project

Summary of Drillhole Length and Drill Platform Number for Collars

Drillhole ID Platform Length Drillhole ID Platform Length

DDB-001 P1 131 DDB-042 P16 176 DDB-002 P1 101 DDB-043 P25 178 DDB-003 P1 133 DDB-044 P16 88 DDB-004 P2 282 DDB-045 P15 83 DDB-005 P2 163 DDB-046 P25 138 DDB-006 P2 300 DDB-047 P15 91


DDB-013 P5' 350 DDB-054 P40 187 DDB-014 P5' 271 DDB-055 P24 197 DDB-015 P5 194 DDB-056 P40 167 DDB-016 P5 223 DDB-057 P23 82 DDB-017 P6 286 DDB-058 P40 600 DDB-018 P6 250 DDB-059 P23 131




DDB-037 P17 173 DDB-078 P37 431 DDB-038 P27 155 DDB-079 P19 304 DDB-039 P17 331 DDB-080 P38 156 DDB-040 P27 155 DDB-081 P19 375 DDB-041 P17 197 DDB-082 P19 186



Table 10.1_2

Berlin Project

Summary of Assay Results from 2010-2011 Drilling Program

Platform Drillhole

ID From (m)

To (m)

Estimated True Width (m)

U3O8 %

U3O8 lb/st

V2O3 (%)

P2O5 (%)

Mo (ppm)

Re (ppm)

Ag (ppm)

Ni (ppm)

Zn (ppm)

Y2O3 (ppm)

Nd2O3 (ppm)

P1

DDB001 109.7 111.2 1.5 0.079 17393 0.68 14.6 294 1.8 5.9 1140 2930 641 154

DDB002 79.2 82.3 3.0 0.137 30185 0.76 9.1 360 5.1 5.9 940 171 700 173

DDB003 80.8 83.8 3.1 0.124 27181 0.71 17.6 626 7.0 3.8 1809 6349 784 160

P2

DDB004 Drillhole did not reach target depth

DDB005 138.7 146.3 7.6 0.152 33426 0.62 10.7 578 8.2 6.2 1462 365 662 177

DDB006 199.7 201.2 1.5 0.046 10098 0.33 5.3 142 1.6 2.3 358 719 273 75

P3

DDB007 152.4 155.5 3.1 0.115 25316 0.45 8.9 632 7.3 2.2 1958 2922 386 85

DDB008 91.4 93.0 1.5 0.068 15057 0.73 14.4 81 0.5 5.5 206 679 916 239

DDB009 94.5 96.0 1.5 0.032 7035 1.07 8.1 63 0.4 8.0 802 2660 260 67

P04

DDB010 207.9 211.9 4.0 0.103 2.26 0.4 7.6 584 5 3 2660 3215 402 105

DDB011 123.2 125.9 2.7 0.126 2.76 0.5 3.7 609 7 3 2197 3319 491 113

DDB012 135.5 138.2 2.7 0.078 1.71 0.3 6.3 404 5 3 1712 2532 340 92

P05 DDB015 161.9 163.9 2.0 0.136 2.99 0.5 10.0 658 7 3 2773 3905 518 124

DDB016 182.2 186.2 4.0 0.099 2.18 0.4 7.6 552 6 2 2178 2920 378 93

P05' DDB013 330.7 332.7 2.0 0.131 2.89 0.5 9.2 741 10 4 4768 4740 554 138

DDB014 259.9 261.9 2.0 0.142 3.13 0.5 10.3 725 8 3 3515 4670 605 150

P06

DDB017 237.5 240.5 3.0 0.091 1.99 0.4 7.7 541 5 3 2580 3333 424 93

DDB018 228.0 229.8 1.8 0.129 2.83 0.4 9.4 558 8 3 2406 3694 608 151

DDB019 225.8 228.4 2.6 0.092 2.03 0.4 7.5 493 5 2 1944 2857 428 102

DDB020 295.6 298.5 3.0 0.091 2.00 0.4 7.4 575 7 3 3240 3293 496 127

P11

DDB031 295.1 297.7 2.5 0.090 1.99 0.3 2.3 409 4 2 1594 2458 368 82

DDB034 296.4 299.4 3.0 0.111 2.44 0.5 2.3 593 6 2 1977 2910 403 86

DDB035 69.9 77.9 8.0 0.120 2.64 0.5 9.7 596 5 3 2434 3237 470 110

DDB035 79.6 82.5 2.9 0.132 2.90 0.5 11.4 619 5 3 2224 3592 446 90

DDB035 83.2 89.9 6.7 0.130 2.87 0.5 9.1 686 6 3 2972 3613 532 118

DDB035 130.7 134.6 3.9 0.162 3.56 0.6 13.0 747 6 4 3056 4230 641 135

DDB035 301.9 304.0 2.1 0.093 2.04 0.4 8.6 671 6 2 2090 2465 290 51

DDB035* 306.6 307.8 Low Recovery

P13

DDB076 74.7 77.6 3.0 0.118 2.59 0.5 9.3 675 6 4 2928 3387 511 118

DDB077 116.3 130.8 3.0 0.149 3.29 0.6 10.6 739 7 3 3120 3807 579 127


P14

DDB062 82.3 83.1 0.8 0.044 0.96 0.3 1.8 122 0 3 738 1600 490 166

DDB062 83.8 84.7 0.9 0.112 2.47 1.0 12.7 23 1 27 1145 3210 1918 508

DDB063 Mineralised layer faulted out


P15 DDB047 39.6 40.7 1.1 0.145 3.19 1.2 14.2 1246 13 4 3411 4894 409 91

P16

DDB042 132.8 135.8 3.0 0.112 2.46 0.4 2.3 563 7 3 2249 3269 467 102

DDB044 65.5 66.1 0.6 0.411 9.03 0.0 21.3 1 0 0 64 226 75 16

DDB044 67.1 68.0 0.9 0.303 6.67 1.1 19.8 1913 19 9 6611 8054 1184 274

P17

DDB037 137.2 138.0 0.8 0.045 0.98 1.0 16.3 196 1 3 1290 2290 1113 312

DDB039 286.5 287.5 1.0 0.042 0.93 0.3 5.3 358 2 2 985 1720 152 33

DDB039 287.9 290.1 2.2 0.159 3.51 0.5 11.3 647 8 4 2806 4295 631 167

DDB041 170.3 171.9 1.6 0.125 2.74 0.8 2.3 226 6 6 3857 3858 589 0

P19

DDB081 352.3 355.0 2.7 0.134 2.96 0.6 11.1 820 8 3 2871 4161 526 111

DDB081 355.1 356.6 1.5 0.060 1.33 0.3 6.6 551 4 4 2406 3768 372 111

DDB082 131.4 132.4 1.0 0.237 5.22 1.1 21.0 1440 21 6 2520 784 926 230

DDB082 132.6 133.7 1.1 0.049 1.07 0.3 7.8 200 4 4 295 100 375 125

P21 DDB067 152.7 154.1 1.4 0.111 2.44 0.5 10.2 418 7 3 1800 3217 507 99 DDB067 154.2 155.3 1.1 0.058 1.27 0.3 7.3 184 2 2 1075 2140 479 136



Table 10.1_2

Berlin Project

Summary of Assay Results from 2010-2011 Drilling Program

Platform Drillhole

ID From (m)

To (m)

Estimated True Width (m)

U3O8 %

U3O8 lb/st

V2O3 (%)

P2O5 (%)

Mo (ppm)

Re (ppm)

Ag (ppm)

Ni (ppm)

Zn (ppm)

Y2O3 (ppm)

Nd2O3 (ppm)

DDB070 253.0 256.4 3.4 0.108 2.38 0.5 8.9 594 5 3 2540 3578 473 116

DDB072 181.4 183.6 2.2 0.098 2.16 0.5 9.1 509 8 3 1999 3324 483 110

DDB072 210.9 211.6 0.7 0.064 1.40 0.4 7.9 443 3 2 1680 1790 282 64 DDB072 212.2 213.0 0.8 0.046 1.02 0.3 6.1 331 2 1 1310 1400 210 51

DDB072 217.9 219.5 1.6 0.090 1.99 0.4 8.9 388 7 3 1485 2422 521 122

DDB074 214.9 215.5 0.7 0.125 2.75 1.3 26.8 49 0 12 1055 2130 1467 297

P23

DDB057 46.3 47.1 0.8 0.089 1.95 1.0 23.5 154 0 2 271 525 1321 328

DDB059 48.8 49.4 0.6 0.405 8.90 1.8 1.3 3850 47 12 6680 2290 411 411

DDB060 76.6 77.6 1.0 0.051 1.13 1.1 11.8 138 0 6 392 1121 762 174

DDB061 88.8 89.4 0.6 0.278 6.13 1.4 26.3 1990 20 5 4930 5280 884 143 DDB061 89.9 90.9 1.0 0.231 5.09 0.9 18.0 1170 13 5 4980 4950 1116 250

P24


DDB048 167.4 168.0 0.6 0.185 4.08 0.6 13.4 809 12 3 2410 4200 572 121 DDB049* 190.5 192.0 Low Recovery

DDB051 241.4 244.0 2.6 0.118 2.59 0.5 8.8 581 7 3 2411 3210 420 103

DDB053 157.0 159.8 2.8 0.084 1.85 0.3 6.7 485 4 2 1751 2541 362 83

P25 DDB046* 104.2 104.7 Low Recovery

P27

DDB036 111.3 115.2 4.0 0.066 1.45 0.3 5.7 533 6 2 1496 2114 246 54

DDB036 161.5 163.5 1.9 0.046 1.00 0.3 4.5 556 5 1 1520 1630 156 31

DDB036 222.9 225.0 2.1 0.092 2.02 0.4 6.8 454 5 2 1643 2492 372 81 DDB038 94.7 97.5 2.8 0.061 1.34 0.4 6.1 490 6 2 1471 1827 208 41

DDB040 100.1 100.8 0.7 0.074 1.64 0.4 7.3 558 6 2 2010 2210 215 44

P28

DDB025 156.3 159.1 2.8 0.133 2.93 0.4 10.0 606 6 3 2743 3787 489 0

DDB026 159.3 162.2 2.9 0.122 2.68 0.4 9.8 630 6 3 2495 3641 443 0 DDB027 160.3 163.2 2.8 0.084 1.84 0.4 7.2 464 4 2 1767 2475 326 0

DDB028 181.0 183.6 2.6 0.104 2.30 0.4 7.7 482 6 2 1990 2986 420 0

DDB029 255.4 258.4 3.0 0.128 2.81 0.5 8.8 632 8 3 3125 3927 526 70

P29 DDB30 Mineralised layer faulted out DDB32 Mineralised layer faulted out

DDB033 370.5 373.2 2.7 0.094 2.06 0.4 7.4 523 4 2 2562 3365 479 115

P30

DDB021 113.2 113.7 0.5 0.070 1.53 0.9 19.0 11 0 1 577 1360 1543 360 DDB022 134.6 135.1 0.5 0.092 2.02 0.8 17.9 128 4 5 2060 2700 1233 324



P37

DDB071 299.1 300.6 1.5 0.140 3.09 0.5 10.4 902 9 3 4278 4451 632 150 DDB073 351.5 353.4 1.9 0.145 3.18 0.6 10.1 1070 12 4 6221 5085 632 146

DDB073 353.6 354.2 0.6 0.182 4.00 0.6 11.7 901 11 6 6280 6080 979 269

DDB075 297.1 298.4 1.4 0.148 3.25 0.5 9.7 772 9 4 3533 4071 644 173 DDB078 404.7 405.1 0.4 0.109 2.39 0.5 11.7 646 5 3 3150 3820 437 79

DDB078 405.6 408.4 2.9 0.153 3.37 0.6 9.5 1079 12 4 6676 5329 699 166

DDB078 410.9 411.3 0.5 0.177 3.89 0.7 11.5 1230 20 6 8810 6930 917 221

DDB078 411.5 412.2 0.7 0.253 5.56 0.8 15.0 1235 14 7 9070 7770 1295 353

P40

DDB056 112.3 114.0 1.7 0.044 0.97 0.4 4.5 351 3 19 1870 2302 165 47

DDB058 Drillhole did not reach target depth

DDB065 245.2 246.2 1.0 0.058 1.27 0.3 5.4 486 4 2 2290 2070 255 51 DDB068 434.2 435.5 1.3 0.151 3.32 0.6 11.2 1270 12 3 6660 5210 635 126

* Intervals with low recovery



11.2 59B59BDiscussion of Drill Results

The geological observations made from the drillhole intersections are reported in Chapter 7, and

are shown in Figures 7.2.2_1 and 7.2.2_2. Due to the limited availability of drilling platforms, the

intersections discussed in Section 10 ranges from true thickness to being oblique to

mineralisation. The bulk of the intersections are expected to be within approximately 30% of the

true thickness of the mineralisation. The mineralisation is defined in more detail in Section 14.

Figures 7.2.2.2_1 and 7.2.2 illustrate the relationship between drillhole orientation and sample

length.

The mineralised intercepts exhibited predominately good recovery, however approximately

25% of the mineralised intervals exhibited low recoveries (~<80%) such that radiometric grade

data had to be used. Of the remaining chemical data used, approximately 90% of the

samples had over 90% recovery recorded.

The overall good geometric continuity of the mineralised region and of the mineralisation tenor

has been described in Chapter 7. Of the 82 drillholes drilled in the 2010-2011 campaign, only

eight did not intersect mineralisation; two holes (DDB-004 and DDB-058) were stopped short

of the mineralised zone and six drillholes on the eastern, overturned limb of the syncline

encountered basement rocks lying directly on overturned black shale strata. These data are

interpreted in terms of thrust faults having removed the mineralised layer in specific areas

creating fault windows. In some cases the drilling is of sufficient density to define the shape

and size of these windows in which mineralisation has been removed. In other areas,

additional drilling is required to adequately define these fault windows.



12 10B10BSAMPLE PREPARATION, ANALYSES AND SECURITY

12.1 60B60BQualification of Personnel and Responsibility for Sampling

All U3O8 Corp. sampling was supervised by qualified, experienced geologists and the actual

samples were taken by appropriately trained geological technicians. All personnel involved with

the sampling in the field and the chain of custody of the samples are employed by, or contracted

by, U3O8 Corp.

All assay data were sent electronically by the analytical laboratory to the Berlin Project Manager

in Colombia, the VP Exploration, whom resides in Argentina, and to the Company’s Technical

Database Manager in Toronto, Canada. QAQC analysis was done by the Technical Database

Manager who responded directly to the laboratory with queries related to the data and

requested reanalysis or whatever remedial actions were required for QAQC purposes.

12.2 61B61BLaboratory Certification

Sample preparation was done in ALS Chemex’s facility in Bogota, Colombia and fire assays

were conducted at ALS Chemex’s facilities in Lima, Peru, while assay by other methods was

done at ALS Chemex’s laboratory in Vancouver, Canada. ALS Chemex is a division of ALS

Minerals which has been in operation for over 60 years and has over 60 analytical laboratories

world-wide. ALS Minerals is certified to ISO 9001 (QC) standards and has an ISO/IEC17025

accreditation from the Standards Council of Canada.

Coffey Mining has reviewed the sample preparation undertaken at the laboratory in Bogota and

concludes that the sample preparation is undertaken to a high industry standard.

12.3 62B62BSampling Procedure

12.3.1 108B108BExploration and Trench Samples

Trench locations and field sampling locations were determined on the basis of radioactivity and

field mapping by the Company’s geologists using standard industry practices. Trenches were

cut perpendicular to strike to expose a cross section of the mineralised unit. The trench face

was mapped to define lithology and to determine its level of radioactivity, which was measured

with Explo-uranium GR-135 and RS-130 spectrometers and / or a Spp2 NF Scintillometer. The

co-ordinates of a reference point located adjacent to the trench were measured by GPS and the

sample positions were recorded with tape and compass measurements from that position,

which was marked with a stake for future reference. Channel sample locations were marked on

the trench face with flagging tape – sample channels were marked perpendicular to bedding to

provide a true-thickness sample. Individual sample length within the channels was determined

to be approximately one metre while respecting geological contacts. Samples were marked

such that one sample does not cut across distinctly different lithologies. All personnel involved

with the sampling wore dosimeters and masks if personnel were entering a deep trench or a

confined space to do the sampling.



Material was removed to a constant depth in the taped channel with a hammer and chisel for

hard rock material or with a spatula or machete in the case of saprolite. Samples typically

consisted of 2-3kg of material. A duplicate sample was taken from each sample site. A

sample number was assigned to each sample from a pre-printed sample book that has two

tear-off tickets that bear the same sample number. One sample number ticket was placed in

the polythene sample bag with the sample, while the other was rolled into the top part of the

bag where it is protected from abrasion. The sample number was copied onto the outside of

the sample bag with a marker pen. The location of the sample and a sample description was

entered into the numbered stub that remains in the sample book. All samples were bagged

and numbered at the location at which they were taken in the field. Sealed samples were

placed in backpacks and were carried by personnel or mules to the Company’s vehicle at the

end of the day. Samples were transported back to the Company’s local office in the Berlin

village in a vehicle that is owned by the Company and was driven by one of its personnel.

The samples were unloaded at the office, ordered, checked and stored in a locked, well

ventilated store room. At approximately weekly intervals, the samples were transported in a

Company vehicle to a town called La Dorada, where the samples were delivered to a national

courier company that provides transport to ALS Chemex’s preparation facility in Bogota.

Duplicate trench samples were stored at the Company’s storage facility at Ibague.

Details regarding location, radioactivity and lithology for each sample were entered into an

Access database.

12.3.2 109B109BDrill Core Samples

Drill core was placed in metal or plastic core boxes by the drill contractor who also marked the

depth at appropriate intervals on the core and on the core box. Rock quality data (“RQD”)

was recorded by technicians at this early stage prior to the core being transported. A

preliminary radiometric log was also done at this time with measurements being taken at

10cm intervals with scintillometers. These data were recorded on field sheets and were then

input into an Access database at the exploration camp in Berlin village.

The core boxes are sealed and carried by mule or aerial ropeway in the early morning and

late afternoon to the nearest road where they were transferred to a waiting Company vehicle.

A Technician accompanied the core from the field to the vehicle. Core boxes were arranged

on logging tables (Figure 11.3.2_1) and the core was ordered and cleaned for detailed review

with core intervals that were marked by the drillers being checked for a second time. RQD

was checked and verified. Detailed geological logging was undertaken on paper forms and

the radioactivity of the core was checked and recorded. The core boxes were then stored in a

locked, well ventilated store room. Every few days, the core was transported from the store

room at Berlin to the Company’s storage facilities in the town of Ibague some 250km from

Berlin. Transport to Ibague was either by Company vehicle or by a commercial transport

company.



At the storage and core handling facility in Ibague, the core boxes were ordered and

geological logs checked by Company geologists. Sample intervals were marked by the

geologists. Sample intervals were typically one metre, but varied according to lithology –

geological contacts of the various lithological units being respected. Radioactivity measured

on the core was compared against, and checked for consistency with, a downhole log of

radioactivity.

The core was then cut with diamond saws in a well ventilated area. The core was cut in half

and then one half was cut again. The ½ core with the two ¼ core segments was returned to

the core box as each section was cut. Sampling then took place with ¼ core being put in

polythene sample bags for assay and the duplicate section of ¼ core being placed in

polythene bags for storage for later use for assay verification or metallurgical testwork, for

example. Both sample bags were numbered using numbered tear-out tags from a sample

book and the drillhole number, depth interval and a brief geological description was entered

into the stub of the sample book and into an Access database. The sample number was

copied onto the outside of the polythene sample bag with a marker pen. Logging and sample

information was recorded on paper forms and this information was then entered into an

Access database.

Duplicate samples and sample blanks were inserted in the sample sequence at pre-

determined intervals; they were numbered such that they were in sequence with mineralised

material. These control samples were similar in appearance to the mineralised material.

Standards were also inserted at pre-determined intervals. In this case, an empty sample bag

was numbered in sequence with the real samples and the appropriate standard, in a sealed

50g sachet, was inserted into the sample bag that was numbered in sequence with the other

samples. QAQC procedures were such that the sample preparation laboratory was requested

to make 10 mesh (“#”, which has a nominal grain size of 2mm) duplicate samples from

mineralised samples at a pre-determined frequency. In these cases, an empty, numbered

sample bag was placed in the sample sequence after the sample from which the duplicate

was to be made. After crushing to 10#, the preparation laboratory split the sample and

inserted one half of the sample into the original sample bag and the other into the empty,

numbered duplicate sample bag.

The sample bags containing ¼ core and QAQC samples for assay were weighed and packed

in boxes for shipment by commercial road transport to ALS Chemex’s sample preparation

facility in Bogota.



Figure 11.3.2_1 Core Sampling and Storage

Drill Rig Setup

Core Transfer by Mule

Core Storage and Mark-up at Berlin

Core Cutting Ibague

Cut Core Showing Sample Intervals

Core Storage at Ibague



12.4 63B63BSample Preparation

On arrival at ALS Chemex’s preparation facility, the following steps were undertaken:

The samples were ordered according to sample number and checked against the sample

submission sheet, assigned a bar-code number and their weight was recorded.

The samples were dried in an oven at 110ºC for eight hours in a stainless steel tray, with

sample numbers attached.

The samples were then jaw-crushed to a target of 70% passing 10#. Size testing was

undertaken approximately every 50 samples.

The 10# material was riffle-split and a sub-sample of 1kg was taken. The coarse reject

was kept by the lab for 45 days before being returned to the client.

The 1kg sub-sample was then pulverised to 85% passing 75µm or better using a LM2

ring and puck pulveriser, for approximately three minutes.

The pulverised sample was then split into a nominal 150g sub-sample and the remainder

of the 10# material was replaced in the original sample bag for storage as a coarse reject

sample.

The ~150g pulp sample was allocated a barcode and boxed for shipment to the assay

laboratory.

ALS Chemex uses commercial transport to send the pulp samples to its assay facilities in

Vancouver, Canada and Lima, Peru.

12.5 64B64BSample Analysis

Different sample analysis methods were required to cover the suite of elements of potentially

economic interest and these analytical procedures are described below under ALS Chemex’s

procedure codes.

12.5.1 110B110BME-MS61U

A 0.25g split of the sample pulp is digested with perchloric, nitric, hydrofluoric and hydrochloric

acids (multi-acid digestion). The residue is topped up with dilute hydrochloric acid and analysed

by Inductively Coupled Plasma - Atomic Emission Spectrometry (ICP-AES). Following this

analysis, the results are reviewed for high concentrations of bismuth, mercury, molybdenum,

silver and tungsten and such samples diluted accordingly. Samples are then analysed by

inductively coupled plasma-mass spectrometry Inductively Coupled Plasma - Mass

Spectrometry (ICP-MS). Results are corrected for spectral inter-element interferences. This

method provides assay results for a 47 element suite, and uses an internal standard certified for

uranium.



12.5.2 111B111BME-M61

This uses the same method as ME-MS61U, yielding the same 47 element suite, but does not

include the internal standard for uranium.

12.5.3 112B112BME-MS81

The decomposition of the sample pulp is done using lithium metaborate fusion (code FUS-LI01)

in which 0.2g of sample pulp is added to 0.9g of lithium metaborate flux and fused in a furnace

at 1,000°C. The resulting melt is then cooled and dissolved in 100mL solution of 4% nitric acid

and 2% hydrochloric acid. This solution is then analysed by ICP-MS. This assay method

provides assay data for 38 elements, including uranium and the full suite of rare earth elements.

12.5.4 113B113BME-XRF

This analytical method is applied specifically to samples that have phosphate content above

the 10% limit provided by the ICP assays. A calcined or ignited sample (0.9g) is added to 9g

of lithium borate flux (50% Li2B4O7- 50% LiBO2), mixed well and fused in an auto fluxer at

temperatures of between 1,050°C and 1,100°C. A flat molten glass disc is prepared from the

resulting melt. This disc is then analysed by X-ray fluorescence spectrometry (“XRF”).

12.5.5 114B114BAA24

This analytical method is fire assay for samples that are periodically assayed for gold. Fire

assay is finished with Atomic Absorption Spectroscopy (“AAS”).

12.6 65B65BComments

Coffey Mining considers that the sampling, preparation, assaying, drilling and storage

procedures undertaken by U3O8 Corp. meet or exceed industry standard practice and are of

high quality and suitable for use in Resource studies. Additional comments on the QAQC are

included in Section 12.4.



13 11B11BDATA VERIFICATION

13.1 66B66BVerification of Data

Coffey Mining conducted a variety of data validation routines to verify the robustness of the

resource database. These activities have included:

Comparison of original laboratory return certificates to the drillhole database.

Comparison of the lithological logging to drillholes observed in the field.

Comparison of the resource database to previous versions of the drillhole database.

The verification checks did not highlight any material issues with the database and the resulting

data was considered appropriate for the use in the following resource estimation study.

13.2 67B67BIndependent Sampling

A total of 12 independent samples were taken by Mr Neil Inwood of Coffey Mining during the

field visit to the Berlin Project from drillholes DDB13, 17 and 20, including three standards.

Under Mr Inwood’s direct supervision the diamond core was cut and bagged.

Once the independent samples were collected and tagged, they were secured with numbered

tamper-proof security tags. The samples were then sent to Coffey Mining’s testing laboratory

in Welshpool, Western Australia for collection.

Upon pick-up from the laboratory in Welshpool, it was noted that the sample bags had been

opened by Australia Customs, and that most samples had been further cut open for closer

inspection. Accordingly, it cannot be absolutely verified that the samples have not been

tampered with, but the author considers this to be unlikely due the high grade nature of the

samples making salting with small particles improbable, the similarity in the multi-element

grade profile, and as most of the samples were reasonably intact upon inspection.

The samples were analysed ICP and XRF by SGS laboratories in Perth, Western Australia.

Analyses for the check and original samples are provided in Table 12.2_1. Overall the samples

gave a similar of result to the original samples, particularly for uranium (2% difference).

13.3 68B68BAnalytical Quality Control Procedure and Data

The Berlin Project QAQC data was supplied to Coffey Mining as a series of spread sheets,

Coffey Mining has checked the veracity of the data presented and considers it to be appropriate.

Samples were taken and submitted for preparation at ALS Chemex (‘ALS’) in Bogota, with

analysis undertaken by ALS in Peru or Canada. Standard ALS Laboratory sample preparation

and analysis procedure was used for analysis of U, V, P, Re, Y, Mo, Ni, and Nd.



Table 12.2_1

Berlin Project

Independent Sampling Results

Drillhole ID

Sample ID From To Length

(m)

U3O8 (ppm) P2O5 (%) V (ppm) Y (ppm) Mo (ppm) Ni (ppm) Ag (ppm) Nd (ppm) Re (ppm)

Original Check Original Check Original Check Original Check Original Check Original Check Original Check Original Check Original Check

DDB13

UC12051 329.68 330.68 1 291 309 3.21 3.01 1040 1270 78.2 100 366 476 1110 1140 1.2 1.3 18.7 22.9 1.81 2.84

UC12052 330.68 331.68 1 1521 1462 10.8 8.29 3420 3070 438 448 870 872 4400 4380 4.4 3.5 100.5 101 10.90 14.00

UC12053 331.68 332.68 1 1155 983 7.7 6.03 2050 2020 434 430 553 555 3670 3430 4.2 3.2 135 128 5.25 6.87

UC12054 332.68 333.23 0.55 248 238 4.47 3.84 806 840 195 216 190 202 539 543 2.5 2.1 63.3 68.6 2.00 2.63

UC12055 333.23 334.23 1 45 52 2.76 2.65 227 222 130 115 33 37 62 74 0.7 0.8 45.2 46.9 0.30 0.40

DDB20

UC12056 294.63 295.58 0.95 235 216 2.7 2.5 1380 1130 72 75.5 185 177 653 561 0.7 0.8 18 18.3 0.96 1.08

UC12057 295.58 296.58 1 466 401 4.94 3.96 1915 1630 148.5 135 492 453 1760 1440 1.1 1.2 33.5 30.5 3.24 3.44

UC12058 296.58 297.58 1 1758 1934 10.55 9.76 3850 3610 676 670 1065 959 6530 6250 4.2 5 182.5 190 12.65 15.90

UC12059 297.58 298.53 0.95 485 444 6.7 5.43 1400 1110 347 313 329 323 1440 1220 2.5 2.9 110 106 3.63 4.10

UC12060 298.53 299.53 1 28 27 2.25 1.82 153 135 109.5 88.2 12 12.8 27 36 0.4 0.4 39.1 36.2 0.09 0.13

DDB17 UC12061 238.4 239.4 1 1475 1568 10.7 8.57 3220 3150 486 478 1010 857 3200 3610 3.3 3.7 105 118 9.58 12.00

UC12062 239.4 240.4 1 826 740 7.42 5.74 1670 1570 399 363 474 425 1540 1690 3.1 3.2 109 116 4.71 5.84

Average 711 698 6 5 1,761 1,646 293 286 465 446 2,078 2,031 2 2 80 82 5 6



The quality control data related to trenching and diamond core drilling has been assessed

statistically using a number of comparative analyses for each dataset. The objectives of these

analyses were to determine relative precision and accuracy levels between various sets of

assay pairs and the quantum of relative error. The results of the statistical analyses are

presented as summary statistics and plots, which include the following:

Thompson and Howarth Plot, showing the mean relative percentage error of grouped

assay pairs across the entire grade range, used to visualise precision levels.

Rank % AMPRD Plot, which ranks all assay pairs in terms of precision levels measured

as the absolute relative difference from the mean of the assay pairs (% AMPRD), used to

visualise relative precision levels and to determine the percentage of the assay pairs

population occurring at a certain precision level. For pulp-based duplicate samples, a

limit of 20% AMPRD is a useful limit to compare and analyse precision from different

datasets. For field duplicates, a limit of 40% AMPRD is a useful limit to compare and

analyse precision from different datasets.

Correlation Plot is a simple plot of the value of assay 1 against assay 2. This plot allows

an overall visualisation of precision and bias over selected grade ranges. Correlation

coefficients are also used.

Quantile-Quantile (Q-Q) Plot is a means where the marginal distributions of two datasets

can be compared. Similar distributions should be noted if the data is unbiased. For

standards and blanks, the Standard Control Plot shows the assay results of a particular

reference standard over time. The results can be compared to the expected value,

providing a good indication of both precision and accuracy over time.

This section will discuss the analysis of the U3O8 Corp. Berlin Project standards, blanks,

duplicates and laboratory standards for both drillholes and trenches.

13.3.1 115B115BU3O8 Corp. Submitted Standards

The following certified standards were used by U3O8 Corp. during their sampling programmes

(Table 12.3.1_1):

ST1000020 expected value of 1,910ppm U3O8

ST1000045 expected value of 1,353ppm U3O8

AMIS-055 expected value of 3206ppm U3O8

Table 12.3.1_1

Berlin Project

U3O8 Corp. Submitted Standards Expected Value for Main Elements

Standard

U (ppm) V (ppm) Mo (ppm) P (%) Y (ppm)

Exp Value

Std Dev

Exp Value

Std Dev

Exp Value

Std Dev

Exp Value

Std Dev

Exp Value

Std Dev

ST1000020 1,910.2 ±96.1 3,414.0 ±105 64.18 ±2.96 1.76 ±0.14

ST1000045 1,353.0 ±65.7 3,584.0 ±176 57.15 ±5.71 2.44 ±0.19

AMIS-055 3,206.0 ±150 53.6 ±5.6



The AMIS standards are manufactured by African Mineral Standards of South Africa; the

ST Standards were created by SGS from material sourced from the Berlin Project. The

U3O8 Corp. standards were assayed for the following elements in addition to U, V, Mo, P, Y, Ni,

Re and Nd.

Uranium (U)

The uranium results are summarised below in Table 12.3.1_2 and summary control plots for the

uranium standards are shown below in Figure 12.3.1_1. Further plots are included in Appendix A.

The U3O8 Corp. submitted uranium standards showed overall good accuracy, although there

is a tendency for a slight negative bias overall. The following comments are made:

The two lower grade standards ST1000020 (EV 1,910ppm U) and ST1000045

(1,353ppm U) show a consistent negative bias of 5.6 and 6.9% respectively.

The higher grade standard AMIS-055 exhibited a positive bias of 4.8%.

Table 12.3.1_2

Berlin Project

U3O8 Corp. submitted U Standards

Standard ST1000020 ST1000045 AMIS-055

Analytical Method ME-MS61U ME-MS61 ME-MS61U ME-MS61U

Element U Result U Result U Result U Result

Units ppm ppm ppm ppm

Detection Limit 1 1 1 1

Expected Value 1,910.20 1,910.20 1,353.00 3,206.00

Expected Value Range 1,719.18 to 2,101.22 1,719.18 to 2,101.22 1,217.70 to 1,488.30 2,885.40 to 3,526.60

Count 45 9 41 14

Minimum 1,540.00 1,760.00 1,190.00 3,200.00

Maximum 2,000.00 1,960.00 1,540.00 3,620.00

Mean 1,802.44 1,848.89 1,259.76 3,360.71

Std Deviation 83.91 66.07 69.23 112.15

% in Tolerance 88.89% 100.00% 82.93% 92.86%

% Bias -5.64% -3.21% -6.89% 4.83%

% RSD 4.66% 3.57% 5.50% 3.34%

Vanadium (V)

The U3O8 Corp. submitted V standards (Table 12.3.1_3 and Figure 12.3.1_2) showed overall

good accuracy and exhibited a slight negative bias. The following comments are made:

Only a small number of standards were available for AMIS055

Standards ST1000020 and ST1000045 show a negative bias varying between 0.6% and

6.2% for V.

Standard ST1000020 have positive bias with one method and negative with another.

Standards ST1000020 and ST1000045 have positive bias varying between 1.9% and

6.4% for P.

Only a small number of standards were available for AMIS-055.

Standard AMIS-055 has positive bias for Y varying between 4.5% and 5.4%.



Figure 12.3.1_1 Berlin Project Uranium Standard Control Plots

1700

1800

1900

2000

2100

2200

15-N

ov-2010

18-N

ov-2010

18-N

ov-2010

02-D

ec-2010

02-D

ec-2010

06-D

ec-2010

16-D

ec-2010

10-Jan

-2011

U_R

esu

lt (

ppm

)

DateReported

Standard Control Plot(Standard ST1000020 ME-MS61 Analysis )

U_Result Expected Value = 1,910.20 EV Range (1,719.18 to 2,101.22) Mean of U_Result = 1,848.89

1500

1600

1700

1800

1900

2000

2100

2200

02-Feb-20

11

09-A

pr-2011

01-Jul-2

011

18-A

ug-2

011

U_R

esu

lt (

ppm

)

DateReported

Standard Control Plot(Standard ST1000020 ME-MS61U Analysis)


1100

1200

1300

1400

1500

1600

17-M

ay-2011

01-Ju

l-2011

30-Ju

l-2011

26-Sep-20

11

U_

Res

ult

(ppm

)

DateReported

Standard Control Plot(Standard ST1000045 ME-MS61U Analysis )


2800

2900

3000

3100

3200

3300

3400

3500

3600

3700

20-Ju

l-2011

30-Ju

l-2011

04-A

ug-2

011

13-A

ug-2

011

18-A

ug-2

011

19-A

ug-2

011

26-A

ug-2

011

05-Sep-20

11

19-Sep-20

11

19-Sep-20

11

19-Sep-20

11

01-O

ct-2011

01-O

ct-2011

U_

Res

ult

(ppm

)

DateReported

Standard Control Plot(Standard AMIS-055 ME-MS61U Analysis )




Figure 12.3.1_2 Berlin Project Vanadium Standard Control Plots

3000

3100

3200

3300

3400

3500

3600

3700

3800

15

-Nov

-201

0

18

-Nov

-201

0

18

-Nov

-201

0

02

-Dec-2

010

02

-Dec-2

010

06

-Dec-2

010

16

-Dec-2

010

10

-Jan

-201

1

V_

Resu

lt (

pp

m)

DateReported

Standard Control Plot(Standard ST1000020 V ME- MS61 Analysis)

V_Result Expected Value = 3,414.00 EV Range (3,072.60 to 3,755.40) Mean of V_Result = 3,274.44

3000

3100

3200

3300

3400

3500

3600

3700

3800

3900

4000

17-M

ay-2011

01-Jul-2

011

30-Jul-2

011

26-Sep-20

11

V_

Res

ult

(pp

m)

DateReported

Standard Control Plot(Standard ST1000045 V ME- MS61U Analysis )


3000

3100

3200

3300

3400

3500

3600

3700

3800

02-Feb-20

11

09-A

pr-2011

01-Jul-2

011

18-A

ug-2

011

V_

Res

ult

(pp

m)

DateReported



290030003100320033003400350036003700380039004000

11-Jan

-2011

28-Feb-20

11

19-A

pr-2011

V_

Res

ult

(pp

m)

DateReported

Standard Control Plot(Standard ST1000020 V ME- MS81Analysis )




Table 12.3.1_3

Berlin Project

U3O8 Corp. submitted V Standards

Standard ST1000020 ST1000045

Analytical Method ME-MS61 ME-MS61U ME-MS81 ME-MS61U ME-MS81

Element V Result V Result V Result V_Result V_Result

Units ppm Ppm ppm ppm ppm

Detection Limit

Expected Value 3,414.00 3,414.00 3,414.00 3,584.00 3,584.00

Expected Value Range 3,072.60 to 3,755.40

3,072.60 to 3,755.40

3,072.60 to 3,755.40

3,225.60 to 3,942.40

3,225.60 to 3,942.40

Count 9.00 44 33 41 9

Minimum 3,100.00 3,040.00 2,950.00 3,020.00 3,320.00

Maximum 3,490.00 3,420.00 3,950.00 3,580.00 4,030.00

Mean 3,274.44 3,203.86 3,393.94 3,348.54 3,740.00

Std Deviation 14127.00% 95.83 256.01 106.49 233.71

% in Tolerance 100.00% 86.36% 75.76% 90.24% 77.78%

% Bias -4.09% -6.16% -0.59% -6.57% 4.35%

% RSD 4.31% 2.99% 7.54% 3.18% 6.25%

Molybdenum (Mo)

The U3O8 Corp. submitted Mo standards (Table 12.3.1_4 and Figure 12.3.1_3) showed overall

good accuracy. Depending on which assay method was used, the results showed a bias

ranging from -10% to + 3.9%.

Table 12.3.1_4

Berlin Project

U3O8 Corp. Submitted Mo Standards

Standard ST1000020

Analytical Method ME-MS61U ME-MS81 ME-MS61

Element Mo Result Mo_Result Mo_Result

Units Ppm ppm ppm

Detection Limit

Expected Value: 64.18 64.18 64.18

Expected Value Range: 57.76 to 70.60 57.76 to 70.60 57.76 to 70.60

Count 41 44 9

Minimum 47.9 52.9 61.2

Maximum 70.5 76.4 70.8

Mean 57.85 65.59 66.71

Std Deviation 4.7 4.21 3.36

% in Tolerance 56.10% 84.09% 88.89%

% Bias -9.86% 2.20% 3.94%

% RSD 8.12% 6.42% 5.03%



Figure 12.3.1_3 Berlin Project Molybdenum Standard Control Plots

40

50

60

70

80

17-May-2011

01-Jul-2011

30-Jul-2011

26-Sep-2011

Mo_

Res

ult

(ppm

)

DateReported

Standard Control Plot(Standard ST1000020 Mo ME-MS61U Analysis)

Mo_Result Expected Value = 64.18EV Range (57.76 to 70.60) Mean of Mo_Result = 57.85

50

60

70

80

02-Feb-2011

09-Apr-2

011

01-Jul-2011

18-Aug-2011

Mo_

Res

ult

(ppm

)

DateReported

Standard Control Plot(Standard ST1000020 Mo ME-MS81 Analysis)


55

60

65

70

75

15-Nov-201

0

18-Nov-201

0

18-Nov-201

0

02-Dec-2010

02-Dec-2010

06-Dec-2010

16-Dec-2010

10-Jan-2011

Mo_

Res

ult

(ppm

)

DateReported

Standard Control Plot(Standard ST1000020 Mo ME-MS61 Analysis )




Phosphorous

The U3O8 Corp. submitted P standards (Table 12.3.1_5 and Figure 12.3.1_4 and 12.3.1_5)

showed overall good accuracy, with a slight positive bias of between 1.8% (ST000020) and 0.8%

(ST1000045).

Table 12.3.1_5

Berlin Project

U3O8 Corp. Submitted P Standards

Standard ST1000020 ST1000045 AMIS-055

Analytical Method XRF12 Me-MS81 XRF12 Me-MS81 ME-MS81

Element P Result P Result P_Result P_Result P_Result

Units % % % % %

Detection Limit

Expected Value 1.76 1.76 2.44 2.44 9.36

Expected Value Range 1.58 to 1.94 1.58 to 1.94 2.20 to 2.68 2.20 to 2.68 8.42 to 10.30

Count 37 33 20 9 8

Minimum 1.8 1.8 2.45 2.45 9.21

Maximum 1.94 1.94 2.54 2.54 9.64

Mean 1.87 1.87 2.49 2.49 9.36

Std Deviation 0.03 0.03 2.00% 0.02 0.12

% in Tolerance 97.30% 96.97% 100.00% 100.00% 100.00%

% Bias 6.36% 6.23% 1.94% 2.10% 0.01%

% RSD 1.75% 1.76% 0.83% 0.94% 1.32%

Yttrium (Y)

The U3O8 Corp. submitted Y standards (Table 12.3.1_6 and Figure 12.3.1_6) showed overall

good accuracy, with a slight positive bias of between 4.5% and 5.4% depending on assay

method for Standard AMIS055. Due to the low number of samples, the results of the analysis

are not robust.

Table 12.3.1_6

Berlin Project

U3O8 Corp. Submitted Y Standards

Standard AMIS-055

Analytical Method ME-MS81 ME-MS61U

Element Y Result Y Result

Units ppm ppm

Detection Limit

Expected Value: 53.6 53.6

Expected Value Range: 48.24 to 58.96 48.24 to 58.96

Count 8 14

Minimum 51.5 51.2

Maximum 58.9 61.4

Mean 55.41 56.29

Std Deviation 2.52 3.01

% in Tolerance 100.00% 71.43%

% Bias 3.38% 5.01%

% RSD 4.54% 5.35%



Figure 12.3.1_4 Berlin Project Phosphorus Standard Control Plots

1.51.61.7

1.81.9

11-Jan-2011

28-Feb-2011

19-Apr-2

011

P_R

esul

t (%

)

DateReported

Standard Control Plot(Standard ST1000020 P ME-MS81 Analysis)

P_Result Expected Value = 1.76EV Range (1.58 to 1.94) Mean of P_Result = 1.87

1.51.61.71.81.9

27-Jan-2011

12-Mar-2011

08-Jun-2011

P_R

esul

t (%

)

DateReported

Standard Control Plot(Standard ST1000020 P ME-XRF12 Analysis )


2.12.22.32.42.52.6

19-Apr-2

011

19-Apr-2

011

07-May-2011

16-May-2011

17-May-2011

08-Jun-2011

08-Jul-2011

18-Jul-2011

20-Jul-2011

P_R

esul

t (%

)

DateReported



2.1

2.2

2.3

2.4

2.5

2.6

23-Feb-20

11

19-A

pr-2011

19-A

pr-2011

19-A

pr-2011

21-A

pr-2011

17-M

ay-2011

26-M

ay-2011

18-Jul-2

011

P_R

esul

t (%

)

DateReported

Standard Control Plot(Standard ST1000045 P ME-MS81 Analysis )

P_Result Expected Value = 2.44 EV Range (2.20 to 2.68) Mean of P_Result = 2.49



Figure 12.3.1_5 Berlin Project Phosphorus Standard Control Plot

Figure 12.3.1_6

Berlin Project Yttrium Standard Control Plots

8.4

8.6

8.8

9.0

9.2

9.4

9.6

9.8

10.0

10.2

10.4

18-A

ug-2

011

19-A

ug-2

011

26-A

ug-2

011

05-Sep-20

11

19-Sep-20

11

19-Sep-20

11

01-O

ct-2011

01-O

ct-2011

P_R

esul

t (%

)

DateReported

Standard Control Plot(Standard AMIS-055 P ME-XRF12 Analysis)


48505254565860

20-Jul-2011

19-Aug-2011

26-Aug-2011

05-Sep-2011

19-Sep-2011

01-Oct-2011

01-Oct-2011

Y_R

esul

t (p

pm)

DateReported

Standard Control Plot(AMIS-055 Y ME-MS81 Analysis )

Y_Result Expected Value = 53.60EV Range (48.24 to 58.96) Mean of Y_Result = 55.41

4850525456586062

20-Jul-2011

30-Jul-2011

04-Aug-2011

13-Aug-2011

18-Aug-2011

19-Aug-2011

26-Aug-2011

05-Sep-2011

19-Sep-2011

19-Sep-2011

19-Sep-2011

01-Oct-2011

01-Oct-2011

Y_R

esul

t (p

pm)

DateReported

Standard Control Plot(AMIS-055 Y ME-MS61U Analysis )

Y_Result Expected Value = 53.60EV Range (48.24 to 58.96) Mean of Y_Result = 56.29



U3O8 Corp. duplicate data samples test the precision (or reproducibility) between two samples

of the same material. As a rule, the precision of the data pairs would increase as the sampling

method moves closer to the pulverisation stage. For the U3O8 Corp. samples, the following

steps would be expected to display increasing levels of precision (from the lowest to highest):

Twin Holes Field Duplicates Coarse reject Duplicate Umpire Pulp Assays

Pulp Repeats Pulp duplicates

U3O8 Corp. undertook the following duplicate QAQC samples:

Field Duplicates.

Blanks.

10 Mesh Duplicates (nominal 2mm grain size).

Blanks

U3O8 Corp. used blanks and these were reported assays below 10ppm with a high of

18.9ppm U. This result is close to detection limit values and indicates no significant

contamination is present during assaying.

13.3.2 116B116BDrillhole Pulp Duplicates

A small number of drillhole duplicate samples were identified. Drillhole duplicate assays for U, V,

Y, Nd, Mo, were within the 10% precision limits and are possibly reasonable. Duplicate assays

for P, Mo and Re showed lower precision levels (Table 12.3.2_1 and Appendix A). Although

there were generally low numbers of data available, the trend is for good reproducibility.

Table 12.3.2_1

Berlin Project

U3O8 Corp. Drillhole Pulp Duplicate Analysis (ALS Lab)

Sample Type

Element Analysis Method

No of Pairs

Mean % HARD

Median% HARD

Within RANK HARD Limit Comparative Means (ppm)

(10%/20%) Original/Duplicate

Drillhole

U ME-MS81 8 2.04 1.47 100/100 318/332

ME-MS61U 15 4.35 2.4 93/93 248/248

V ME-MS81 8 2.25 0.77 100/100 1,899/1,939

ME-MS61U 15 2.56 1.55 100/100 1,179/1,171

Mo ME-MS81 8 3.49 2.6 88/100 371/171

ME-MS61U 15 5.98 3.08 80/100 199/199

Y ME-MS81 8 2.78 1.5 100/100 1118/119

ME-MS61U 15 2.78 1.16 93/100 102/103

Nd ME-MS81 8 1.83 1.36 100/100 33.78/33.64

Ni

ME-MS61U

15 2.54 2.2 100/100 664/659

P 15 5.5 1.15 87/93 5,287/5,117

Ag 15 1.53 1.12 93/100 1.77/1.79

Re 15 5.48 3.23 87/93 1.75/1.76

A small number of coarse crush duplicate results were available, further results are required

to allow for an adequate assessment of the coarse-crush precision levels.



Coffey recommends that more pulp duplicate samples be submitted for assaying, and that

duplicate samples also be taken from the coarse crush to assess the precision of the riffle

splitting.

Trench Pulp Duplicates

U3O8 Corp. submitted trench duplicates showed good precision, although Ni showed slightly

lower precision levels than the other elements (Table 12.3.2_2). Summary control plots are in

Appendix A.

Table 12.3.2_2

Berlin Project

U3O8 Corp. Trench Pulp Duplicate Analysis (ALS Lab)

Sample Type


No of Pairs

Mean % HARD

Median% HARD



Trenches

U

ME-MS81

12 1.14 0.98 100/100 219/217

V 12 0.96 0.94 100/100 2,237/2,231

Y 12 0.94 0.72 100/100 222/222

Nd 12 0.98 0.85 100/100 77/77

Mo 12 0.95 0.83 100/100 70/70

Ni 12 4.34 2.3 83/100 91/94

Lab Blanks and Pulp Duplicates

The laboratory also undertook the blanks and pulp duplicate samples as part of the internal

quality control process. The source material for the internal blanks is not known and these

internal blanks are not considered to be of commercial quality and depending on the quality of

mixing during presentation, could produce disparate results. On this basis the internal blanks

have been excluded in the analysis.

The lab pulp duplicate analysis for all elements shows very good RANK HARD precision for

the few sample pairs analysed (Table 12.3.2_3).

Table 12.3.2_3

Berlin Project

Summary of Data Precision U308Corp Lab Pulp Duplicate Analysis (ALS Lab)

Sample Type


No of Pairs

Mean % HARD

Median% HARD



Lab Duplicates

U

ME-MS81

7 1.3 1.33 100/100 378/378

V 10 1.99 1.19 100/100 2,889/2,876

P 10 2.15 0.64 100/100 783/764

Y 10 0.98 0.87 100/100 342/343

Ni 10 1.91 1.42 100/100 87/88



13.4 69B69BComments

The QAQC data available for the Berlin Project showed overall good accuracy and precision.

The following comments and recommendations are made:

The U3O8 Corp. assaying shows good levels of accuracy and precision, and the

resulting assay database is suitable for use in resource estimation studies.

Lower grade U standards should also be sourced (e.g. 200ppm U, 500ppm U, 800ppm U)

so as to test the accuracy of the lower-level U assays.

Additional umpire and coarse-crush duplicates are required to allow for an analysis of the

coarse-crush precision levels.

It is recommended that standards, blanks, pulp duplicates and Umpire pulp duplicates be

sampled at a rate of 1:20.



14 12B12BMINERAL PROCESSING AND METALLURGICAL TESTING

14.1 70B70BApproach

Preliminary metallurgical work by Minatome showed recoveries of approximately 85% on

uranium and 50% on vanadium, albeit with an elevated acid consumption of about

350kg/tonne of ore (Roussemet & Houot, 1979). Minatome’s work did not attempt to extract

other metals.

The challenge at Berlin is the carbonate-rich nature of the majority of the host rock that leads

to high acid consumption. Baseline leach tests were done on crushed, whole ore

(not beneficiated) to establish metal extraction and reagent consumption. These baseline

tests, which included acid and alkaline leach, were done by SGS Lakefield in Ontario and the

Australian Nuclear Science and Technology Organization (“ANSTO”) near Melbourne in

Australia. Given the high acid consumption for moderate recoveries obtained from acid leach

and alkaline leach only being effective for a small component of the suite of metals of potential

economic interest, work at ANSTO was suspended until a means of beneficiation could be

found for the mineralised material. ANSTO made a referral for the beneficiation work to be

undertaken at Optimet in South Australia, and this work is underway.

A suite of initial flotation tests had already commenced at SGS Lakefield’s facility in Ontario

and it was decided to continue with that work under the direction of Mr John Goode.

Ferric leach tests were undertaken by SGS OreTest in Perth, Australia, under the guidance of

Dr Paul Miller, the principal of Sulphide Research Processing Pty Ltd. Given the results

obtained from the ferric leach, SGS Lakefield was asked to conduct duplicate tests on different

composite samples than the two used by SGS OreTest as a check on the initial results.

Testwork continues at SGS OreTest and SGS Lakefield.

14.2 71B71BLaboratories and Consultants

The metallurgical testwork reported below was undertaken at the following laboratories:

SGS Lakefield OreTest Pty Ltd in Perth, Australia. SGS Lakefield OreTest was established

as a metallurgical services company in 1993 as Lakefield OreTest Pty Limited and is now a

subsidiary of the SGS Lakefield group, which has been offering mineral processing

services to the mining industry since 1948;

SGS Lakefield in Ontario, and predecessor companies, have been undertaking metallurgical

testwork for over 50 years and its Lakefield facility is ISO/IEC 17025 accredited;

ANSTO was formed in 1987. It is a State agency within the portfolio of the Commonwealth

Department of Innovation, Industry, Science and Research. ANSTO is responsible for

delivering specialised advice, scientific services and products to government, industry,

academia and other research organisations. It does so through the development of new

knowledge, delivery of quality services and support for business opportunities.



Metallurgical testwork on Berlin was carried out under the guidance of the following consultants:

Mr John Goode, P.Eng., a Qualified Person within the definition of that term in NI 43-101 of

the Canadian Securities Administrators, has overseen the metallurgical testwork carried out

by SGS Lakefield. Mr John Goode P.Eng., is a metallurgist with 49 years experience, much

of it in the uranium industry. He is a member of the Canadian Institute of Mining, Metallurgy

and Petroleum and is a fellow of the Australasian Institute of Mining and Metallurgy.

Dr Paul Miller, a Qualified Person within the definition of that term in NI 43-101 of the

Canadian Securities Administrators, has overseen the metallurgical testwork carried out by

SGS Lakefield OreTest, and verified the technical information relating to the tests reported

from that laboratory in this report. Dr Miller is a metallurgist who has specialised in

hydrometallurgy and has over 30 years’ experience in the commercial application of

processes for the treatment of sulphide-bearing ore. Dr Miller has a doctorate in Chemical

Engineering, is a member of the Institute of Mining and Metallurgy, London, and is also a

Chartered Engineer. He is currently Managing Director of Sulphide Resource Processing

Pty Ltd. Dr Miller is responsible for the design and interpretation of the ferric leach tests

discussed in this Section 13.

Information relating to the drillhole intersections used in the composite samples is provided by

Dr Richard Spencer, President and CEO of U3O8 Corp., and a Qualified Person within the

definition of that term in NI 43-101 of the Canadian Securities Administrators.

14.3 72B72BNature of Material

The metallurgical testwork was carried out on quarter core or sample reject (1/2 core samples

that were jaw crushed to ~2mm grain size). Of the 82 drillholes drilled in the 2010-2011 drill

campaign at Berlin, 74 intersected the mineralised horizon. Material from 25 (34%) of these

intersections have been used in the various metallurgical tests (Table 13.3_1). The

intersections used in the metallurgical testwork cover a wide geographic area within the area

drilled in the 2010-2011 campaign (Figure 13.3_1) and hence are considered to be

representative of the mineralisation encountered to date at Berlin.

Individual samples with sample numbers that correspond to the samples that were used for

assay, were shipped to the various laboratories where the samples were crushed to 10#

(~2mm diameter), blended together and homogenised to constitute the composite samples

listed in Table 13.3_1.



Table 13.3_1

Berlin Project

Details of Drillhole Intercepts Used in Metallurgical Testwork to Date

Composite Sample Lab Material Facies Drillhole Number From (m)

To (m)

Sample Type Mass (kg)

Colombian Comp

SGS Lakefield

Weathered, saprolite-like material

Sandstone & Siltstone DDB-001 109.7 111.2

¼ core 3

DDB-002 77.7 85.1 5

DDB5

Rock from drill core

Sandstone DDB-005 138.65 144.76 ¼ core 9

DDB7 Limestone DDB-007 152.42 157.16 7

Berlin Comp Sandstone & Limestone 2kg of Comp DDB5 and 2kg of DDB7 were mixed ¼ core 4

DDB10-15 Limestone

DDB-10 207.89 211.89

10# & ¼ core 42

DDB-011 123.2 125.94

DDB-012 135.46 138.2

DDB-013 328.68 332.68

DDB-014 257.98 261.9

DDB-015 161.14 163.86

UC Comp SGS OreTest Rock from drill core Limestone

DDB-018 226.27 229.77

10# & ¼ core 11

DDB-019 225.78 227.78

DDB-020 294.63 298.53

DDB-025 155.44 159.09

DDB-026 158.42 161.32

BER-160611

ANSTO Rock from drill core Limestone

DDB-016 181.45 186.95

¼ core 39

DDB-017 236.86 240.94

DDB-018 226.27 229.77

DDB-019 225.22 228.35

DDB-020 294.63 298.53

DDB-025 155.44 159.09

DDB-026 158.42 162.22

DDB27-38 or ANSTO 2

DDB-027 160.32 163.16

¼ or ½ core 49

DDB-028 180.95 183.55

DDB-029 253.79 258.93

DDB-031 295.13 297.66

DDB-033 370.45 373.15

DDB-034 296.35 300.3

DDB-036 111.25 115.2

222.93 225.0

DDB-038 94.69 97.53

287.91 290.14



Figure 13.3_1. Location of Intercepts Used in Metallurgical Testwork on the Berlin Project Relative to the

Rock-Type Encountered in Drilling

14.4 73B73BHead Grade & Composition of Composite Samples

The head grade of the various samples used in metallurgical testwork is shown in Table 13.4_1.

There are two main geological facies or mineralisation types at Berlin: one is a sandstone facies

that has a relatively low carbonate content (composite samples Colombian Comp and DDB5),

while over 90% of the uranium resource is hosted in carbonate facies (composite samples

DDB7, DDB10-15, BER-160611, UC Comp and DDB27-38, Figure 13.3_1 and Table 13.4_1).

One composite, sample Berlin Comp, is a mixture of material from DDB5 and DDB 7 and is a

mixture between sandstone and carbonate mineralised material.

Estimated mineral content of the ANSTO 1 composite sample is shown in Table 7.3.3_1. The

principal acid consuming minerals are carbonate, organic carbon and phosphate.



Table 13.4_1

Berlin Project

Head Grade of Composite Samples Used in Metallurgical Testwork by the Various Laboratories

Labs SGS Lakefield ANSTO SGS OreTest

Element Units

Composite Sample Names

Heads Colombian

Comp

DDB7 Comp

DDB5 Comp

Berlin Comp

DDB10-15 Comp

ANSTO 1 or BER-160611

UC Comp

ANSTO 2 or DDB27-38

U3O8 %

U % 0.045 0.079 0.130 0.130 0.085 0.085

Re ppm 2.7 5.6 11.7 -- -- 4.67 5.19

CO2 % 1.69 25.60 0.99 9.82 26.90

S % 1.04 0.63 0.92 -- 0.86 0.24 0.97 0.94

SiO2 % -- -- -- 45.00 14.40 8.60 7.50 4.00

Ag ppm 2.3 1.7 5.9 5.4 < 4 2.3 2.4

Al % 2.92 0.66 3.09 2.22 0.67 0.74 0.68 0.70

As ppm 80 101 212 97 169 130 205 145

Ba ppm 1,050 1,080 1,900 1,310 957 850 880 1,729

Be ppm 2.3 < 3 < 3 < 2 < 0.8 0.6 0.7

Bi ppm < 20 < 20 < 20 < 0.6 < 20 0.3 0.2

Ca % 4.49 28.60 13.10 16.08 30.95 28.50 27.20 31.60

Cd ppm 11 19 13 15 27 20.7 24

Co ppm 8 5 10 6 < 10 9 6

Cr ppm 317 287 774 547 274 700 868 507

Cu ppm 150 80 184 131 85 157 103

Fe % 2.85 0.47 1.17 1.43 0.50 0.53 0.81 0.73

K % 0.74 0.22 0.78 0.56 0.21 0.24 0.24 0.26

Li ppm < 20 < 20 < 20 18 < 10 7 10

Mg % 0.41 0.32 0.42 0.38 0.49 0.43 0.39 0.51

Mn ppm 39 50 41 77 155 158 155

Mo ppm 343 455 777 496 501 531 402

Na % 0.04 0.03 0.05 0.04 0.07 <0.01 0.04 0.04

Ni ppm 739 1,830 2,010 1,350 2,420 2,485 2,151

P % 2.17 3.17 5.94 3.78 3.36 2.9 3.23 3.50

Pb ppm < 80 < 30 119.00 48.00 < 30 68 14

Sb ppm 56 54 67 63 65 51 47

Se ppm 192 202 601 350 285 232 180

Sn ppm < 20 < 20 < 20 2.00 < 20 0.9 0.4

Sr ppm 486 618 924 663 736 838 805

Ti % 0.22 0.05 0.14 0.13 0.04 0.038 428 489

Tl ppm < 60 < 30 < 30 15.40 < 30 7.5 6.6

V % 0.30 0.21 0.50 0.34 0.25 0.19 0.23 0.25

Zn % 0.06 0.21 0.06 0.11 0.28 0.26 0.25

Au ppm 0.04 0.02 0.02 -- --

Ce ppm 85 59 154 119 66 73 110

Dy ppm 17 17 48 31 22 21 26

Er ppm 13 13 36 21 16 16 19

Eu ppm 3 3 9 6 4 4 5

Gd ppm 17 17 51 32 23 24 29

Ho ppm 4 4 12 7 6 5 6

La ppm 128 135 392 240 171 186 263

Lu ppm 2 2 5 3 2 2 2

Nd ppm 73 67 196 128 89 80 89

Pr ppm 18 16 47 32 21 21 31

Sc ppm 7 3 7 6 3 4 3

Sm ppm 13 12 35 21 16 15 22

Tb ppm 3 3 7 5 3 3 4

Th ppm 9 3 9 9 3 <0.01 3 <.05

Tm ppm 2 2 5 3 2 2 2

U ppm 449 -- -- -- 853 929 735

Y ppm 225 230 664 415 326 320 363 340

Yb ppm 11 10 29 17 13 12 13

Zr ppm 0 0 0 0 0 400 11 6



14.5 74B74BMetallurgical Testwork

14.5.1 117B117BBaseline Leach Tests on Raw (Unbeneficiated) Mineralised Material

Alkaline Leach

ANSTO tests BER 3A and 4B were done under aggressive alkaline leach conditions with a low

slurry density of 2% solids and at temperatures of 90° and 250°C and reagent concentrations as

shown in Table 13.5.1_1. Recoveries for uranium were not significantly different (67% and 68%

extraction) under the very different temperature conditions (250°C and 90°C respectively). The

recovery of vanadium, phosphate and nickel under alkaline conditions was poor. Molybdenum

recovery was good at 91%, especially under the 90°C conditions.

Acid Leach

Baseline sulphuric and hydrochloric acid leach tests were done to determine rates of extraction

of uranium and other metals under various conditions as shown in Table 13.5.1_2. Recoveries

for uranium were moderate to good, but with high to very high levels of acid consumption.

In all sulphuric acid leach tests, sodium chlorate was added to achieve a target Oxidation-

Reduction Potential (ORP) of 500mV when measured using a platinum electrode against a

Ag•AgCl saturated KCl reference electrode.

14.5.2 118B118BAcetic Acid Pre-Leach Followed by Strong Acid Leach

SGS Lakefield Test 32 used acetic acid as a means of dissolving calcite preferentially without

dissolving significant amounts of apatite as described by Gharabaghi et al. (2010). Experimental

work by these authors had shown that, under certain leach conditions, organic acids selectively

attack calcite over apatite. Acetic acid reacts with calcite to form calcium acetate from which

acetic acid can be regenerated by the addition of sulphuric acid which results in the precipitation

of calcium sulphate (gypsum) as follows:

CH3COOH + CaCO3 CaCOOH + CO2 (1)

CaCOOH + H2SO4 CaSO4 ↓ + CH3COOH (2)

Test 32 was an acetic acid test on raw mineralised material (Composite DDB10-15) ground to

100µm. The test was done over 3.5 hours with 20% acetic acid under the conditions shown in

Table 13.5.2_1.

Approximately 35% of the carbonaceous gangue was removed by using acetic acid. This

approximation is based on a comparison of the calcium content of the feed (30.9%) and the

residue (27.3%) and factoring in a 29% mass loss between the feed and residue (411g

original sample mass reduced to 290g in the residue). Analysis of the solution showed that it

contained only 4.7% of the uranium and very little phosphate (8ppm P).



Table 13.5.1_1

Berlin Project

Composite Sample BER-16061 - Summary of Alkaline Leach Conditions and Metal Extractions Achieved

Test No.

Conditions Extractions (%)

Time Hrs

Grind p80µm

Temp °C

pH Pulp Density

% Solids

Na2CO3 Addn kg/t

NaHCO3

Addn kg/t

Oxidant Addn kg/t

U V P Y Mo Ni

BER 3A 6 35 250 N/R 2 4,000 1,000 0 66.7 22.0 0.0 N/R 88.1 0.0

BER 46 48 12 90 10 2 2,000 500 16.2 68.2 27.2 3.4 N/R 91.1 11.3

Table 13.5.1_2

Berlin Project

Various Composite Samples - Summary of Acid Leach Conditions and Metal Extractions Achieved

Laboratory Lab

Test # Sample

Grind P80 (µm)

T (°C)

Pulp Density

(%)

Leach Time

(Hours)

Average pH

Acid Metal Recovery (%)

Type Acid Added

(kg/t)

Net Acid Consumption

(kg/t) U V P Y Mo Ni Re

SGS

11 Colombian Comp 41 80 50 48 0.94 H2SO4 194 158 70.8 51 31 67 15 19

13 DDB7 46 65 50 48 2.54 H2SO4 712 652 78.0 72 52 34 32 12

14 DDB5 38 65 50 48 1.61 H2SO4 225 183 59.8 45 49 20 29 6

16 DDB5 15 50 25 4 0.95 HCl 175 104 45.5 9 62 50 17

17 Berlin Comp 70 50 25 4 0.95 HCl 214 193 33.3 5 25 27 10

32 DDB10-15 100 20 2 3.5 5.1 Acetic 385 4.7 1 <1

ANSTO BER 1A BER-160611 35 60 2 24 1.5 H2SO4 477 468 76.8 24.6 28.1 43 40

BER 1B 35 60 2 24 1.5 H2SO4 144 139 68.3 29.5 14.9 66 75

SGS OreTest AJ1033 UC Comp 100 25 10 48 2 H2SO4 558 550 19 10.9 18.7 16.9 11 30

Grind P80 (μm) = 80% passing size in micrometres.

Sodium chlorate addition in SGS Lakefield sulphuric acid leach tests were 4 to 7kg/tonne.



Table 13.5.2_1

Berlin Project

Summary of Acetic Acid Pre-Leach Conditions and Subsequent Strong Acid Tests and Metal Extractions Achieved

Laboratory Lab Test

# Sample

Grind P80 (µm)

T (°C)

Pulp Density

(%)

Average pH

Acid Metal Recovery (%)

Type Concentration

(g/L) Acid added

(kg/t)

Net Acid Consumption

(kg/t) U V P

SGS

32 DDB10-15

100

20 33 5.1 Acetic 385 4.7 1

55 Test 32 low-carbonate Residue

81 25 3.69 H2SO4 50 715 595 95.3 84 74

56 81 25 2.16 HCl 50 605 432 96.2 87 98

Grind P80 (μm) = 80% passing size in micrometres.

Sodium chlorate addition in Test 55 was 23kg/t of leach feed or 16kg/t of original material.

V and P extractions are approximate



The low-calcite residue from Test 32 was leached with a solution of 50g/L sulphuric acid (Test 55)

and 50g/L hydrochloric acid (Test 56). Total uranium extractions were 95% with sulphuric acid

and 96% with hydrochloric. Sulphuric acid addition in Test 55 was 715kg/t residue, which is the

equivalent of 505kg/t on an original ore mass basis allowing for the 25% mass loss in Test 32.

Hydrochloric acid addition in Test 56 was 605kg/t residue, which is the equivalent of 427kg/t on

an original ore mass basis. A baseline leach test was not done on DDB10-15 but the CO2

content is very similar to that for DDB7 which was leached with sulphuric acid in Test 13 and

consumed 712kg/t suggesting a 29% reduction in sulphuric acid demand. Phosphorus extraction

was 59% from the sulphuric acid leach and 71% from the hydrochloric acid leach.

Results of the acetic acid test confirm that the organic acid does dissolve calcite preferentially,

consuming approximately 35% of the carbonaceous gangue while dissolving a negligible

amount of apatite. Uranium recoveries after the acetic acid pre-leach are significantly higher

than those obtained by direct sulphuric acid leach on similar composite samples (95%

recovery in Test 55 (acetic acid pre-leach) compared with 78% in Test 13 (Table 13.5.1_2 and

Table 13.5.2_1). Acetic acid can be regenerated with the acidification of the calcium acetate

– and the efficiency of this process will be tested on the Berlin ore. These preliminary tests on

the use of organic acids to preferentially dissolve carbonaceous material show promise for

incorporation in the processing of Berlin ore.

14.5.3 119B119BFerric Leach

SGS OreTest Testwork

The work has demonstrated that relatively mild leaching conditions can be applied to the

treatment of Berlin material for extraction of a broad variety of elements. Two Samples of ground

ore from the Berlin Project, described as ”UC Composite” and ”ANSTO 2”, were subjected to mild

acidic ferric leaching at atmospheric pressure and a temperature of 65°C for 48 hours at SGS

OreTest’s facility in Perth. Residues from testing were subjected to “re-leaching” using 10%

solutions of either hydrochloric acid or sulphuric acid. The average extractions of elements from

a total of nine batch tests are shown in Table 13.5.3_1. Good mass balances were obtained from

the tests conducted.

The leach parameters investigated were particle size, pulp density, temperature and ferric iron

concentration.

Particle Size

Although three different particle sizes were investigated (106µm, 75µm, 38µm), there is an

insignificant change in metal recoveries and some of the superior results were obtained from

those tests using the larger size range of 106µm (Table 13.5.3_1). Although more extensive

tests are required to verify the influence of particle size, these results are promising in

suggesting that an acceptable grind size associated with normal milling operations may be

sufficient to liberate elements and expose them sufficiently for leaching.

UC Comp and ANSTO 2 are composite samples made from 18 drillhole intercepts (~24% of

all intercepts drilled in the 2010-2011 drill campaign).



Table 13.5.3_1

Berlin Project

Summary of Ferric Leach Tests Undertaken by SGS OreTest on Two Composite Samples From Berlin

Composite Sample #

Maximum Grain Size

(µm)

Ferric Leach (Stage1) Releach / Wash (Stage 2) Elemental Extraction %

Concentration (g/L)

Pulp Density

(%)

Temp. (°C)

Acid Type & Conc.

Pulp Density

(%)

Temp. (°C)

Uranium Vanadium Phosphate Yttrium Neodymium Zinc Nickel Molybdenum Rhenium

UC Comp

106 50 10

65

10%HCL 10

40

98.4 73.2 99.2 94.8 51.0 99.4 61.7 48.0 33.8

75 50 5 10%HCL 10 97.2 79.6 93.6 96.1 89.5 64.3 61.1 58.9 49.6

38 25 5 10%HCL 10 98.8 82.0 99.5 95.7 94.7 98.1 59.7 47.7 29.7

38 50 10 10%HCL 10 96.6 74.6 99.1 94.6 86.4 99.9 50.1 43.4 58.0

Average recovery from Composite Sample # 1 with hydrochloric acid wash 97.8 77.4 97.9 95.3 80.4 90.4 58.2 49.5 42.8

ANSTO 2 106 50 5

65 10%HCL 10

40 97.2 80.2 99.5 95.8 82.8 98.5 62.7 56.8 27.1

106 50 10 10%HCL 10 96.4 78.9 92.5 95.7 84.6 98.6 77.4 61.1 71.2

Average recovery from Composite Sample # 2 with hydrochloric acid wash 96.8 79.6 96.0 95.8 83.7 98.5 70.1 59.0 49.2

Average recovery from Composite Samples UC Comp & ANSTO 2 with hydrochloric acid wash 97.3 78.5 96.9 95.5 82.0 94.5 64.1 54.2 46.0

UC Comp 106 50 10

65

10%H2SO4 10

40

93.1 56.0 97.9 79.6 48.9 93.7 58.6 44.8 22.8

ANSTO 2 106 50 5 10%H2SO4 10 97.3 69.4 99.4 89.8 63.6 97.0 62.7 53.8 11.1

106 50 10 10%H2SO4 10 98.0 73.3 99.4 89.0 66.4 97.0 76.4 55.7 64.6

Average recovery from Composite Samples UC Comp & ANSTO 2 with sulphuric acid wash 96.1 66.3 98.9 86.1 59.6 95.9 65.9 51.4 32.8

Average recovery from all 9 tests on Composite Samples UC Comp & ANSTO 2 97.1 75.2 97.6 93.1 76.2 94.2 63.6 52.7 42.2



Temperature

The influence of temperature was evaluated in two ferric leach tests that were done at 40°C and

65°C (Table 13.5.3_2). The higher temperature, 65°C test, resulted in improved recoveries for

most metals. All tests were conducted at atmospheric pressure and no noxious gases or fumes

were generated.

Leach Stages

Although the testwork consisted of two stages of leaching, the results suggest that the major

influence of the second stage was to improve recovery by making soluble those elements

which had been leached in the first stage but had co-precipitated with iron species.

Pulp Density

Low pulp densities of 5% and 10% were chosen in order to ensure a high ratio of reagent

liquid to ore in Step 1, creating conditions that are favourable for effective leaching. As pulp

densities rise, leach efficiencies may be expected to decrease moderately. This effect was

tested in the doubling of the pulp density from 5% to 10% and results show no significant

change in recoveries. This positive result suggests that ferric leach is likely to be efficient at

higher pulp densities approaching those typically used in commercial operations (40-50%),

which would reduce the amount of reagent liquid required and would lower costs. Further

work will be conducted to optimise pulp density.

Ferric Concentration

Ferric iron concentrations of either 25g/L or 50g/L were tested at SGS OreTest with both

achieving comparable metal recoveries. These data suggest that there is potential to reduce

ferric iron consumption by using higher slurry densities.

Corroborative Testwork by SGS Lakefield

Given the success of the ferric leach testwork undertaken by SGS OreTest in Perth, Australia, a

second, corroborative test was undertaken by SGS Lakefield in Ontario using different

composite samples from Berlin. Composite DDB10-15 is from six hole intersections from

carbonate ore facies and DDB5 is from sandstone ore facies (Figure 13.3_1) as a check as to

the application of ferric leach to a different, albeit minor, ore type. With the original testwork

done by SGS OreTest and that done by SGS Lakefield, core from 25 intersections out of

74 intersections (34%) drilled in the 2010-2011 campaign have undergone ferric leach testwork.

The conditions used for the corroborative test were similar to those used in the original

testwork and showed similar uranium recoveries (Table 13.5.3_3). It can be noted that the

residue for Test 47 is anomalously high and is being checked.



Table 13.5.3_2

Berlin Project

Summary of Ferric Leach Tests Conditions (Without Stage 2 Leach) Undertaken by SGS OreTest Showing Effect of Temperature on Metal Recoveries

Composite Sample #

Maximum Grain Size

(µm)

Concentration (g/L)

Pulp Density

(%)

Temp. (°C)

Final pH (after 48hrs)

Elemental Extraction %

Uranium Vanadium Phosphate Yttrium Neodymium Zinc Nickel Molybdenum Rhenium

UC Comp 106 50 10 40 1.9 50.9 2.5 1.8 9.6 nd nd 34.2 7.0 nd

106 50 10 65 2.3 65.4 4.7 5.9 17.7 nd nd 55.2 4.1 nd

nd = not determined

Table 13.5.3_3

Berlin Project

Uranium Extraction Obtained in Corroborative Ferric Leach Tests Undertaken by SGS Lakefield in Ontario

Composite Sample

Ferric Leach (Stage 1) Re-Leach (Stage 2) Combined Stage 1 and 2 Extraction (%) Leach Test Number

Pulp Density (% solids)

Time (h)

Uranium Extraction (%)

Leach Test Number

Acid Type and Conc.

Uranium Extraction (%)

DDB10-15

39 10

24 79.8 43 10% H2SO4 89.5 97.9

48 80.7 47 10% H2SO4 31.2 86.7

51 10% HCl 90.8 98.2

40 5

24 82.1 44 10% H2SO4 79.6 96.3

48 87.8 48 10% H2SO4 62.8 95.5

52 10% HCl 89.6 98.7

DDB5

41 10

24 85.5 45 10% H2SO4 66.8 95.2

48 71.8 49 10% H2SO4 78.8 94.0

53 10% HCl 76.0 93.2

42 5

24 70.9 46 10% H2SO4 96.3 98.9

48 85.4 50 10% H2SO4 70.6 95.7

54 10% HCl 71.5 95.8

Average Uranium Extraction with 24h Primary Leach and Sulphuric Acid Re-leach 97.1

Average Uranium Extraction with 48h Primary Leach and Sulphuric Acid Re-leach 93.0

Average Uranium Extraction with 48h Primary Leach and Hydrochloric Acid Re-leach 96.5

Notes: Grind size was 75 micrometres

Primary leach samples were taken after 24h and used to determine extraction and used in re-leach

Primary leach temperature was 65°C

Re-leach was at 10% solids and temperature was 40°C



The Role of Ferric Iron

One of the most significant features of the ferric leach tests is that no acid additions were

made to control pH. The acid required for this first stage of leaching was provided by means

of the natural acidity of the ferric sulphate solution (pH 1.5 to 2.5) combined with the

generation of sulphuric acid by the natural precipitation of ferric iron as hydroxide. This is

illustrated in the reaction given below in which each mole of ferric sulphate generates between

two and three moles of acid:

Fe2(SO4)3 + 6H2O ↔ 2Fe(OH)3 + 3H2SO4

Ferric sulphate Water Ferric iron hydroxide Sulphuric acid

The second stage re-leach used only a relatively mild acid strength of either 10% sulphuric

acid or hydrochloric acid. Although the tests were conducted at a low pulp density, the

principal role of the acid supplied is to act as a carrier for the ferric iron which is the dominant

reactant for leaching. This is well illustrated by comparison of metal recoveries from the acid

leach tests (Table 13.5.1_2) with those of the ferric leach (Table 13.5.3_1 and 13.5.3_2) that

shows higher rates of extraction for uranium as well as a broad suite of elements in the ferric

leach tests.

Reagent Consumption

Additional work needs to be done to define the consumption of ferric iron more precisely – its

consumption is complicated by the dynamic reactions through which ferric changes to ferrous

iron and vice versa.

The fate of the ferric lixiviant used at 50g/L can be determined based on the reactions most

likely to occur for carbonate neutralization and phosphate release. Theoretically, there are

four possible iron species or end products which could form under the test conditions used:

Ferric is un-reacted and remains in solution as ferric sulphate Fe2 (SO4)3;

Ferric is hydrolysed to form the solid goethite FeO.OH (this constitutes a lesser

component at the temperature of 65°C used in the current tests);

Ferric is hydrolysed to form the solid ferric hydroxide Fe(OH)3 (this constitutes the major

component in current tests); and

Ferric is reduced to ferrous sulphate (FeSO4) which is soluble.

Using sensible reaction assumptions the fate of ferric lixiviant from Test 39 (10% pulp density,

75µm grain size, 65°C temperature and a concentration of 50g/L ferric iron) appears to be as

follows (Table 13.5.3_4):



Approximately 4% of the ferric iron remains un-reacted and is still available for reaction or

recycling in the liquor;

Up to 19% of the ferric appears to have formed goethite solid which will be lost as it is

largely resistant to dissolution with sulphuric acid. It could be dissolved with hydrochloric

acid. However, goethite does not appear to capture soluble uranium and therefore its

dissolution is not necessary;

Up to 68% of the ferric iron appears to have formed ferric hydroxide solid which can be

recovered in a re-leach step and re-used. Dissolution of the hydroxide is necessary in

order to remobilise some uranium that co-precipitated;

Approximately 9% of the ferric iron has been converted to ferrous sulphate in solution

and is available to be reconverted to ferric and cycled.

The above inventory suggests that up to 19% of the iron has been permanently lost from the

system as insoluble goethite-like precipitates. This presumes that re-leaching of the residue

to remobilise ferric iron is carried out together with re-conversion of ferrous to ferric iron. This

also makes the assumption that acid is available for the re-leach and re-conversion stages. If

insufficient acid is available for reconversion, then further iron is lost from the system. In the

current test scenario, the ferric make-up requirements would be 84kg of ferric per tonne of ore

due to the iron lost as goethite (presuming sulphuric acid and not HCl is used for re-leaching).

This goethite has contributed 241kg/tonne of acid to the overall acid needs.

No external acid additions were made in Test 39 (the ferric leach step) and the reaction of ferric

to give the different iron end products created a minimum of 740kg/tonne of acid for reaction.

This value is in reasonable agreement with the theoretical requirement of approximately

766kg/tonne of acid for neutralization of carbonates and release of phosphate.

In the ferric leaching step, approximately 80.7% of the uranium was extracted and remained in

the ferric liquor. On acid washing of the ferric leach residue (Test 43), the uranium extraction

was increased to an overall value of 97.8% by solubilising uranium which had co-precipitated

with iron. The measured amount of acid consumed in the acid washing step was 125kg/tonne

of ore equivalent and all the iron remobilised was present in the ferric form. Allowing for the

need to convert the soluble ferrous iron formed from ferric leaching back to ferric iron for re-

use, a further 34kg/t acid would be required. Therefore, the total acid consumption would be

159kg/t with a ferric consumption of 84kg/t.

Further work is currently underway to define more precisely the mechanisms of the reactions

in order to optimise the leach process. Specifically, the balance between the precipitation of

iron that generates acid versus addition of acid to allow more ferric iron to be recirculated into

the leach solution requires thorough definition.



Table 13.5.3_4

Berlin Project

Calculation of Reagent Consumption in the Two-Step Ferric Leach followed by Acid Wash Tests Undertaken at SGS Lakefield

Test ID Sample Feed

Density (%)

Acid Target, Conc Mass

Increase Fe Lixiviant Unreacted

Fe Lixiivant pptated

Fe Converted to Ferrous Solution

Fe Re-Mobolised

in Acid Re-Leach

Fe Lost (neither

mobilised nor as Ferrous)

Initial ferric Lixiviant Inventory

kg/t

Kg/t Ferric Lixiviant Make-up Required

Kg acid / t provided by Fe pptation and

Ferrous Generation

Kg Acid / t Consumed in Acid Releach

Kg Acid / t to be Consumed Reconverting Ferrous Back

to Ferric

Total Overall kg/t Acid Make-up for Releach

+ Ferrous Conversion

% Uranium Extracted

FERRIC LEACH AND RELEACH TESTS DDB10-15 Comp 10% solids

Test 39 DDB10-15 Comp 10 50g/L Fe3+ added as ferric sulphate 84.0% 3.9% 87.4% 8.7% 18.6% 450.0 83.7 1070.2 124.6 34.4 159.0

80.7%

Test 43 Releach Test 39 RES 24h 10 10% Sulphuric for Re-leach -20.7% 68.8% 89.5%

%overall wt increase 46.0% %U Total = 98.0%


80.7%



Test 39 DDB10-15 Comp 10 50g/L Fe3+ added as ferric sulphate 84.0% 3.9% 87.4% 8.7% 65.3% 450.0 294.0 1070.2 0.0 34.4 34.4 80.7%


80.7%

Test 51 Releach Test 39 RES 48h 10 10% HCl Hydrochloric for Re-leach -67.9% 62.1% 90.8%

%overall wt increase -41.0% %U Total = 98.2%

FERRIC LEACH AND RELEACH TESTS DDB10-15 Comp 5% solids

Test 40 DDB10-15 Comp 5 50 g/L Fe3+ added as ferric sulphate 101.0% 37.4% 28.9% 7.7% 0.8% 950.0 7.6 786.5 654.1 64.1 718.2

87.8%




87.8%




87.8%

Test 52 Releach Test 40 RES 48h 10 10% HCl Hydrochloric for Re-leach 61.1% 26.5% 89.6%


FERRIC LEACH DDB5 Comp 10% solids

Test 41 DDB5 Comp 10 50g/L Fe3+ added as ferric sulphate 0.1% 73.4% 8.1% 18.5% 1.1% 450.0 4.9 169.4 124.0 73.0 197.1

71.8%




71.8%




71.8%

Test 53 Releach Test 41 RES 48h 10 10% HCl Hydrochloric for Re-leach -37.1% 7.5% 76.0%


FERRIC LEACH DDB5 Comp 5% solids

Test 42 DDB5 Comp 5 50g/L Fe3+ added as ferric sulphate -5.5% 64.1% 26.4% 9.4% 2.1% 950.0 19.7 739.5 N/A 78.6 N/A

85.4%



Test 42 DDB5 Comp 5 50g/L Fe3+ added as ferric sulphate -5.5% 64.1% 26.4% 9.4% 0.3% 950.0 3.1 739.5 138.6 78.6 217.2

85.4%



Test 42 DDB5 Comp 5 50g/L Fe3+ added as ferric sulphate -5.5% 64.1% 26.4% 9.4% 0.3% 950.0 3.0 739.5 33.3 78.6 111.9

85.4%

Test 54 Rleach Test 42 RES 48 h 10 10% HCl Hydrochloric for Re-leach -32.3% 26.1% 71.5%




14.5.4 120B120BFlotation Tests

Berlin Composite Sample

Six scoping flotation tests were carried out on the “Berlin Comp” composite sample as follows:

Five of the tests focused on removing apatite from the mineralised material:

The two initial tests (F1 and F2) were conducted on whole ore and examined pre-

flotation of fine, organic carbon ahead of an apatite flotation circuit;

The following three tests were completed on stage-ground and deslimed feed that

went to apatite flotation (Tests SB1-SB3).

The sixth test (Test SB4) was undertaken on stage-ground and deslimed feed and

attempted to float uranium minerals.

Results of the flotation tests can be summarised as follows:

F1 - Desliming in F1 was unsuccessful because of the fineness of the grind (80% passing

50µm). Pre-flotation to remove the graphitic carbon was not successful, recovering ~10%

of the uranium in ~7% of the mass, but it was clear that carbon was still floating in the

phosphate roughing stages, suggesting that higher frother dosages or a frother blend could

improve the recovery of graphitic carbon and U3O8 to the pre-float concentrate.

F2 – Pre-flotation of the graphite yielded results similar to those of F1 with 9% of the

uranium in ~7% of the mass.

Tests SB1 to SB4 used stage-grinding to avoid overproduction of fines. Desliming of the

stage-ground ore resulted in recovery of 27%-30% of the U3O8 in ~14% of the mass.

Attempts to selectively recover apatite in tests SB1 to SB3 were unsuccessful due to the

large amount of graphitic carbon in the ore and possibly due to poor liberation. The same

conclusion can be made for test SB4, where it was attempted to selectively float the

uranium minerals.

DDB10-15 Composite Sample

Given that the graphite in the prior tests appeared to inhibit flotation, all of the subsequent

tests on composite DDB10-15 included a carbon pre-flotation circuit. This composite sample

also has a higher carbonate content than Berlin Comp on which the prior flotation tests were

done. Six tests were done on the DDB10-15 composite sample as follows (Table 13.5.4_1):

DDB-F1 & DDB-F2: Tests DDB-F1 and DDB-F2 were designed to investigate the effect

of grind size on apatite flotation. The tests included a pre-float carbon step, followed by

apatite flotation. Test DDB-F1 was done at a relatively coarse grain size of 104µm and

DDB-F2 on a finer grain size of 58µm.

DDB-F3 & DDB-F4: In these tests, Step 1 was carbon pre-leach, Step 2 was a sulphide

flotation stage using PAX (a sulphide collector), Step 3 was a dilute hydrochloric acid

leach of the sulphide rougher tails prior to apatite flotation in order to clean the calcite

surfaces in an attempt to reduce accidental activation.



DDB-F5 & DDB-F6: Step 1: A charge of composite DDB10-15 was ground to a K80 of

~60µm. Step 2 involved the use of acetic acid to dissolve as much calcite as possible. The

sample was leached with 220kg/t acetic acid for two hours and filtered. Step 3 was carbon

flotation and Step 4 was apatite flotation in the case of DDB-F5 and calcite flotation in the

case of DDB-F6.

Results:

DDB-F1 and DDB-F2 results showed that carbon recovery in the pre-flotation step, which

yielded 16% mass recovery, was better at the finer grains size; 64% at 58µm and 51% at

104µm. Selectivity between apatite and calcite was poor with 13% of the calcite and 15%

of the apatite reporting to the carbon concentrate (both tests) and 30%-31% of the uranium.

Leach tests were conducted on the DDB-F1 and DDB-F2 carbon pre-float material as well

as on the apatite rougher concentrates and results are as follows:

DDB-F1 Prefloat leach (Leach Test 68) – from the 31% of the uranium that reported to

the carbon pre-float concentrate (16% of the sample mass), uranium extraction was

91% with a sulphuric acid consumption of approximately 467kg/tonne of feed or

73kg/tonne of ore (Table 13.5.4_2).

DDB-F2 Prefloat leach (Leach Test 69) – 30% of the uranium reported to the carbon

pre-float concentrate (16% of the sample mass), from which uranium recovery peaked

at 85% at 48 hours but dropped to 77% at 72 hours, with an acid consumption of

605kg/tonne of feed or approximately 99kg/tonne of ore (Table 13.5.4_2).

An acid leach test (Test 64) was done on the apatite concentrate from DDB-F1 that

contained about 43% of the sample mass and 52% of the uranium. Uranium

extraction was 97% and vanadium recovery was 92% with acid consumption of


An acid leach test (Test 66) was done on the apatite concentrate from DDB-F2 that

contained about 51% of the sample mass and 55% of the uranium. Uranium

extraction was 96% and vanadium recovery was 90% with acid consumption of


Leach extraction kinetics in the above tests were rapid and leach times of between 24

and 48 hours would possibly be optimal (Figure 13.5.4_1).

DDB-F3 & DDB-F4 results showed that 49%-52% of the sulphides were recovered in the

carbon pre-float and together, the carbon pre-float and sulphide circuit recovered 87-88%

of the organic carbon with 70%-71% of the uranium, 55%-57% of the phosphate and 48%-

59% of the calcite in 47% of the mass.

DDB-F5 & DDB-F6 results show that the acetic acid leach dissolved 26%-28% of the

calcite with only 3%-4% of the uranium. In the pre-float rougher step, 77%-82% of the total

organic carbon was recovered with 49%- 56% of the uranium, 39%-47% of the calcite and

24%-26% of the apatite in 26%-35% of the sample mass. Addition of a cleaning stage

reduced the mass of the carbon concentrate to 13% of the sample (DDB-F5) or 18.1%

(DDB-F6) containing 40% of the uranium (both tests), 29% (Test DDB-F5) to 32% (Test

DDB-F6) of the calcite and 25% (both tests) of the apatite.



Table 13.5.4_1

Berlin Project

Summary of Carbon and Sulphide Pre-float and Apatite or Carbonate Flotation Tests Results by SGS Lakefield on Composite Samples “Colombian Comp” & DDB10-15

Test No Flotation Product

Mass % Distribution

% U3O8 Total Organic

Carbon Calcite Apatite Dolomite Sulphur

F1

Slimes 0.4 0.6 0.6 0.4 0.3 0.5

Pre-Float 6.4 9.6 14.3 4.5 4.6 7.6

Rougher Concentrate 61.3 67.5 68.5 68.3 68.6 65.8

Rougher Tail 31.9 22.3 16.6 26.9 26.5 26.1

Head 100.0 100.0 100.0 100.0 100.0 100.0

F2

Pre-Float Conc 6.8 9.2 12.9 5.3 5.7 8.2

Rougher Concentrate 26.8 42.5 39.7 22.1 29.2 32.3

Rougher Tail 66.4 48.3 47.4 72.6 65.1 59.5 Head 100.0 100.0 100.0 100.0 100.0 100.0

SB-1

Slimes 14.0 29.7 10.8 9.8 20.4

Rougher Concentrate 11.8 17.9 11.1 19.9 12.1

Rougher Tails 74.2 52.4 78.1 70.3 67.5

Head 100.0 100.0 0.0 100.0 100.0 100.0

SB-2

Slimes 14.0 26.8 11.1 9.6 21.2

Rougher Concentrate 16.9 21.8 16.3 22.0 21.2

Rougher Tails 69.1 51.4 72.6 68.4 57.7

Head 100.0 100.0 0.0 100.0 100.0 100.0

SB-3

Slimes 14.0 28.0 11.5 9.5 21.5 Rougher Concentrate 11.6 14.0 9.6 21.7 14.2

Rougher Tails 74.4 58.0 78.9 68.8 64.3

Head 100.0 100.0 0.0 100.0 100.0 100.0

SB-4

Slimes 14.0 27.5 10.9 9.5 20.4

Rougher Concentrate 31.8 37.0 28.9 41.2 34.2 Rougher Tails 54.2 35.5 60.1 49.2 45.4

Head 100.0 100.0 0.0 100.0 100.0 100.0

DDB-F1

Pre-Float Conc 15.6 30.7 51.1 12.8 15.4 15.1 Apatite Float 43.3 51.9 32.2 43.5 56.6 42.7

Rougher Tail 41.1 17.4 16.7 43.7 28.0 42.2

Total 100.0 100.0 100.0 100.0 100.0 100.0

DDB-F2

Pre-Float Conc 16.3 30.4 63.6 12.7 14.9 15.9

Apatite Float 51.0 55.3 26.3 53.0 60.4 50.3 Rougher Tail 32.7 14.3 10.1 34.2 24.7 33.8

Total 100.0 100.0 100.0 100.0 100.0 100.0

DDB-F3

Pre-Float Conc 26.9 45.1 72.1 23.0 28.8 26.0 48.7

Sulphide Conc 20.1 26.2 14.4 17.9 28.6 19.5 35.3

Apatite Float 19.8 14.7 3.6 22.5 24.4 20.0 5.2

Rougher Tail 33.2 14.0 9.9 36.6 18.2 34.5 10.8

Total 100.0 100.0 100.0 100.0 100.0 100.0 100.0

DDB-F4

Pre-Float Conc 29.2 47.9 76.4 24.7 31.7 27.7 52.1

Sulphide Conc 17.4 21.8 11.8 15.9 23.6 16.9 30.2

Apatite Float 8.9 7.1 3.0 9.2 9.7 9.5 8.6

Rougher Tail 44.5 23.2 8.8 50.2 35.1 45.9 9.0

Total 100.0 100.0 100.0 100.0 100.0 100.0 100.0

DDB-F5

Pre-Float Cleaner Conc 12.8 39.6 67.2 8.7 17.5 13.1

Pre-float Cleaner Tail 13.4 16.8 9.7 15.3 19.0 16.2 Apatite Float 24.6 21.1 10.0 33.4 35.7 28.8

Rougher Tail 33.7 19.4 13.0 42.6 27.8 42.0

PLS 15.5 3.1 0.0 0.0 0.0 0.0

Total 100.0 100.0 100.0 100.0 100.0 100.0

DDB-F6

Pre-Float Cleaner Conc 18.1 39.6 73.5 10.1 23.8 17.4 Pre-float Cleaner Tail 16.7 9.5 8.0 18.5 20.6 18.8

Apatite Float 41.5 39.2 16.3 39.4 51.4 46.9

Rougher Tail 9.4 8.1 2.2 4.4 4.2 8.0 PLS 14.4 3.7 0.0 27.6 0.0 8.9

Total 100.0 100.0 100.0 100.0 100.0 100.0



Table 13.5.4_2

Berlin Project

Summary of Sulphuric Acid Leach Test Results Undertaken at a Temperature of 65°C on Flotation Products by SGS Lakefield on Composite Sample DDB10-15

Leach Test Sample

% Distribution in Sample Extraction from Leach Feed Sulphuric Acid

(kg per tonne of Feed) Sulphuric Acid Consumption per Tonne of Ore

(kg/tonne) Mass (%)

U3O8

(%) U3O8

(%) V2O5

(%) Added Consumed

Test 68 DDB-F1 Carbon Pre-float 15.6 30.7 90.65 80.4 605 467 73

Test 64 DDB-F1 Apatite Concentrate 43.3 51.9 96.75 92.4 835 702 304

Test 69 DDB-F2 Carbon Pre-float 16.3 30.4 76.96 76.5 727 605 99

Test 66 DDB-F2 Apatite Concentrate 51.0 55.3 96.40 90.4 849 725 370

Grind - as received Feed density - 33% solids Acid target - 50g/L Temperature target - 60°C EMF target - 500mV Leach time - 72 h



Figure 13.5.4_1 Leach Kinetics on Flotation Products from Composite Sample DDB10-15

Mineralogy indicates that uranium is associated with carbonaceous material in the deposit, as

is observed in petrographic studies detailed in Section 7. The association is evident in the

flotation data where a carbon pre-float is able to concentrate the uranium. The linear nature

of the relationship is illustrated in Figure 13.5.4_2.

Figure 13.5.4_2

Plot Of Total Organic Carbon Recovery in Pre-Float Concentrate Versus Uranium Recovery

‐

10

20

30

40

50

60

70

80

90

100

0 10 20 30 40 50 60 70 80

U extracted (%)

Leach time (h)

Berlin flotation productsUranium leach kinetics

DDBF1 C Prefloat

DDBF1 Con

DDBF2 C Prefloat

DDBF2 Con



14.6 75B75BSummary and Conclusions

Baseline testwork confirmed the work of Minatome that heated agitated sulphuric or

hydrochloric acid leach was moderately effective at extracting uranium from whole ore.

Leach extractions ranged from 19% to 78%, but acid consumption was high at 100 to

650kg/tonne of ore. Aggressive alkaline leach achieved uranium recoveries of 67%-68%,

but with poor recoveries for associated metals and phosphate, with the exception of

molybdenum, from which recoveries of 88%-91% were achieved.

A two-step process involving ferric leach followed by acid wash of the sample residue

obtained outstanding recoveries for a wide range of elements of potential economic

interest at Berlin. The principal features of the ferric and acid wash process are:

The ferric leach was done on whole ore, without prior beneficiation. Tests were

done on three different grind sizes and no significant difference in recovery related

to grain size was noted. These data suggest that a grind to approximately 100µm is

adequate for effective ferric leaching.

Ferric leach was done at a relatively low temperature of 65°C at atmospheric pressure.

Uranium extractions from ferric leach followed by hydrochloric wash ranged from 96%-

99% and 93%-98% with sulphuric acid wash. Initial tests were done at SGS OreTest

on 24% of the mineralised intersections cut in the 2010-2011 drill campaign at Berlin.

Corroborative testwork was undertaken by SGS Lakefield and uranium recoveries

using hydrochloric acid wash were 93%-99%. Using a sulphuric acid wash, uranium

recoveries were 87%-98% or 94%-98% if the anomalous result for Test 47 is omitted

pending confirmation by reassay. The corroborative tests were from different intercepts

from those used by SGS OreTest and hence ferric leach testwork has now been

conducted on 34% of the intercepts cut in the 2010-2011 drill program. Given the high

percentage of intercepts used in the testwork and the wide distribution of the samples

through the area drilled, the ferric leach testwork is representative of the part of the

Berlin Project drilled in the 2010-2011 campaign, on which the resource estimate is

based.

This two-step leach process provides good to excellent recoveries for phosphate (94%-

99%) and excellent to acceptable recovery for a range of metals of potential economic

interest including Vanadium (56%-82%), Yttrium (80%-96%), Neodymium (49%-95%),

Zinc (64%-100%), Nickel (50%-77%), Molybdenum (43%-61%) and Rhenium (11%-

71%).

No acid was added in the first step of the leach process. The ferric iron generates

sufficient acid to neutralise the carbonate and phosphate. Initial estimates that take

into account the extent to which iron species can be recycled back to ferric iron for re-

use in the leach process, ferric consumption is approximately 84kg/tonne of ore and

sulphuric acid consumption is approximately 159kg/tonne of ore.

This testwork indicates that there is potential to decrease reagent consumption through

the use of higher pulp densities and lower ferric iron concentration in solution. Further

testwork is being undertaken to confirm these conclusions.



Having established the efficiency of the two-step ferric leach followed by acid wash, any

means of increasing the efficiency of the process or reducing potential reagent input

costs would benefit the economics of the project. To this end, flotation tests and initial

testwork with acetic acid was undertaken as follows:

Flotation tests:

Showed good potential to extract a sulphide concentrate for further processing to

extract these metals. Approximately half of the sulphides (nickel sulphides,

sphalerite and pyrite) are currently recovered in the carbon pre-float. The use of a

sulphide collector in a subsequent step increased sulphide recovery to 82%-84%.

Pre-flotation tests results suggest that uranium is associated with organic carbon.

This conclusion is supported by petrographic evidence.

A carbon pre-float step provides a relatively efficient means of separating organic

carbon (graphite) from the principal acid-consuming species: apatite, calcite and

dolomite. Test DDB-F2, for example, shows that 64% of the carbon is

concentrated in 16% of the sample mass with 13% of the calcite, 15% of the

apatite and 16% of the dolomite. 30% of the uranium also reports to the carbon

pre-float. Acid leach tests on the graphite pre-float concentrate shows recoveries

of 85%-91% after 48 hours with a sulphuric acid consumption of 73-99kg/tonne of

ore.

Further testwork showed that up to 88% of the graphite could be removed from

the ore by pre-flotation, with 70%-71% of the uranium. However, the apatite and

calcite content of the pre-float, graphitic material also increases.

Extraction of the majority of the graphite may improve efficiencies of separation of

calcite and apatite by flotation.

The calcite/U3O8 ratio in the ore used in the DDB series of flotation tests was 617.

In the best of the flotation tests (DDB-F3), the calcite/U3O8 ratio was reduced to

337 suggesting that acid consumption per unit of uranium would be roughly

halved by flotation. Additional tests could lead to improvement in uranium

recovery and the calcite/U3O8 ratio.

Acetic acid tests:

An acetic acid pre-leach removed approximately 35% of the calcite while

extracting less than 5% of the uranium and reducing the solids weight by 29.4%.

The residue was successfully leached in both sulphuric acid and hydrochloric acid

with acid addition of 715kg/tonne H2SO4 or 605kg/tonne HCl on a residue basis,

calculated to be 505kg/tonne and 427kg/tonne respectively on a dry ore basis.

Uranium extraction for these tests was 95% (sulphuric) and 96% (hydrochloric).

Acetic acid leaching prior to carbon pre-flotation flotation (DDB-F5 and DDB-

F6) removed approximately 26% of the calcite with only 3-4% U3O8 losses.

The carbon pre-float cleaner concentrate on the acetic acid leach residue

recovered 70% of the graphite, 40% of the U3O8, and 11% CaO in ~15% of the

mass. Hence, acetic acid was effective as a pre-leach step.



The results achieved in the metallurgical testwork undertaken to date have identified key

areas for further work and optimization as follows:

There is potential to increase the efficiency of sulphide flotation.

A carbon pre-float step shows promise as a means of extracting the majority of the

graphite and associated uranium. Graphite tends to interfere with the separation of

calcite from apatite and its removal may aid in their separation. Further tests should

attempt to maximise extraction of graphite in as small a volume as possible with a

relatively high uranium content and minimum apatite content. Flotation would then

attempt to maximise the separation of apatite from calcite to reduce reagent consumption

in the subsequent leach process.

There is potential to use smaller amounts of acetic acid to simply roughen, or otherwise

modify, calcite surfaces relative to apatite as a means of potentially improving the

efficiency of separation by flotation. Having established the selectivity of acetic acid for

calcite through this testwork, the focus should be on refining the parameters that allow

the acid to etch the calcite while minimizing acetic acid consumption. This may be

achieved through very dilute, or short duration, acetic acid leach than those tested to

date. If this approach is effective, regeneration of acetic acid with sulphuric acid addition

should be tested and refined.

Ferric leach with subsequent acid wash has proved to be an excellent means of recovery

for not only uranium, but also an extensive suite of potentially economic metals and

phosphate. Additional work needs to be done to more fully define reagent consumption

and the extent to which the ferric iron and sulphuric acid can be regenerated and recycled.

Testwork needs to investigate higher pulp densities and lower ferric iron concentrations.

Pyrite occurs in the hanging wall of the mineralisation and in the Cretaceous sequence and

also in the adjacent country rocks. Local pyrite may provide a cost-effective source of ferric

reagent and/or sulphuric acid.

As a suitable beneficiation and leach process becomes better defined, preliminary

investigation is required into solvent extraction or ion exchange processes for the recovery

of uranium and associated metals. Tailings and effluent treatment tests should also be

undertaken to obtain knowledge of the behaviour of the final leach residue and barren

solutions.



15 13B13BMINERAL RESOURCE ESTIMATES

In January 2012 a maiden resource estimate was estimated for the Berlin Project. The

Qualified Person responsible for the Berlin resource estimate was Mr Neil Inwood, who is a

Principal Resource Geologist with the consultancy Coffey Mining Pty Ltd. The Qualified

Persons’ certificate for Mr Inwood is included in Appendix B. The details of the resource

estimations are summarised in the following section.

15.1 76B76BResource Database and Validation

15.1.1 121B121BDatabase

The resource database comprised 82 diamond drillholes for 18,523m, drilled from 29 platforms.

Up to six drillholes were drilled from each platform in an essentially east-west fan orientation

with dips ranging from 42° to 90°. Three of the holes were orientated to the north and two

towards the south from platforms P21, P23 and P40. The Berlin estimate used 79 drillholes for

a total of 313 samples. A combination of chemical assaying (216 samples – 69% of the total)

and radiometric eU3O8 data (99 0.8m composites) was used to estimate the mineralised zone.

Where there was no or poor core recovery (typically <70%), the uranium grade was estimated

on the basis of radiometric data (eU3O8). No adjustment of the radiometric eU3O8 grade data

was considered to be required as comparative analysis of chemical assaying versus radiometric

values showed a good relationship. The ranked assays used for the resource estimate were

assigned to the resc_u3o8 field in the database.

A total of 38 trenches have also been dug for the Berlin Project up to 8m in length and

averaging 3.5m for a total of 132m. The trench assays were not included in the estimate, but

were used in aiding the interpretation of the mineralised zone near surface.

Drilling coverage for the project area is quite varied because of limitations of suitable drilling

platforms but ranges from 60m x 100m up to 200m x 300m on the northern and southern

extents.

15.1.2 122B122BValidation

The 2011 drillhole database was checked using the following methods:

Drillhole collars against the supplied surface topography.

Hole depths for the geology log, survey log and assay intervals didn’t exceed the hole

depth.

That sample IDs and grade data retuned from the laboratory match the data in the

database.

That valid codes e.g. lithology, geotechnical log etc. have been used.

Sampling intervals were checked for gaps and overlaps, and against previous versions of

the database.

3D continuity of the mineralisation.



The following additional checks were also undertaken prior to estimation:

Overlapping intervals.

Missing intervals.

Checks of the top ~300 assay intervals to the original digital laboratory files.

3D analysis of collar positions and downhole survey traces.

No significant validation errors were detected in the database and the database was

considered appropriate for the use in the resource estimation.

15.2 77B77BGeological Interpretation and Modelling

15.2.1 123B123BGeological and Mineralisation Model

Uranium and associated mineralisation at Berlin is contained within a phosphorous-rich

limestone layer within a north-north-west to north trending syncline. Typically, little significant

mineralisation occurs in the surrounding sediments. The mineralised layer ranges between

0.6m and 7m thick and averages 3m. Most of the drillholes along the eastern margin intersect

the mineralisation twice. A number of drillholes with downhole intercepts >8m along the eastern

margin have been interpreted as intersecting down the length of the eastern fold hinge, in

particular holes DDB-035 and DDB-077 and therefore do not represent true thickness.

Evidence for the synclinal structure was also supported by 3D structural measurements of the

diamond drill core.

The mineralisation constraints were generated based upon sectional interpretation and three

dimensional analysis of the available drilling and trench information at the surface. The

mineralised zone (Figure 14.2.1_1) was modelled as one contiguous 3km zone, trending north-

north-west to north and plunging approximately 10° to the north using a nominal lower cutoff

grade of 400ppm U3O8. Figure 14.2.1_2 shows a typical sectional interpretation of the

mineralised layer.

15.3 78B78BWeathering and Topographic Profile

The topographic surface for the Berlin Project area was defined based upon a supplied DEM

surface. The project has insufficient data to determine the depth of the weathering profile so

to account for weathering the surface topography was dropped 10m. Further work is required

to understand and model the weathering profile.



Figure 14.2.1_1 Modelled Mineralisation Used for 2012 Resource

Showing mineralised outline, drilling and trench locations



Figure 14.2.1_2 Oblique Section 618000mN with Drilling and U3O8 Values



15.4 79B79BStatistical Analysis

15.4.1 124B124BRadiometric Data

Radiometric eU3O8 grade data was sourced from the Mount Sopris Probe used by Gaia Energy

(Colombia) Ltd. Comparative intervals with good recovery (43 in total) of eU3O8 grade and

chemical grade data were compared statistically and visually along section. The statistical

analysis indicated that there was no overall bias present between the two datasets with the

mean of the radiometric data being 590ppm eU3O8 and the mean of the chemical data being

600ppm U3O8. Figure 14.4.1_1 shows a scatter plot of the two datasets.

Figure 14.4.1_1

Scatter Plot of Chemical and Radiometric U3O8 Data

Although globally the datasets are similar, there are examples of local variance and further

analysis is recommended on a larger dataset. Chemical assaying is the preferred method for

this deposit.

15.4.2 125B125BSummary Statistics and Top Cuts

After analysis of the radiometric data, a ranked U3O8 field was used to populate the data for

the uranium grade estimate. Chemical assays were preferentially used, however for intervals

of low recovery, and thus low confidence in the grade data, radiometric grades were used.



The average sample interval was 0.85m; accordingly 0.8m composites were used to generate

the grade data for the estimate. Intervals with missing data were ignored. A statistical analysis

was carried out on the 0.8m composite data to determine an appropriate top cut to apply to the

data. The approach taken included:

Review of the 3D grade distribution;

Review of the histogram and probability plots with significant breaks in populations used

to identify possible outliers;

Ranking of the individual composites and investigating the effect of the higher grades

upon the standard deviation and the mean of the data population.

Analysis of the data determined that top-cutting the multi-element assay data was not necessary

for the estimates. Table 14.4.2_1 shows summary statistics of the 0.8m multi-element

composite data. Figure 14.4.2_1 shows histogram plots of the 0.8m composite data from within

the mineralised zone. Figures 14.4.2_2 and 14.4.2_3 show scatter plots of U3O8 against the

multi-element data.

15.4.3 126B126BBulk Density Data

Density data was based upon 27 measurements from within mineralised material

(>100ppm U3O8) using the core diameter/volume method. This data was used as only four

density readings were available from within the more tightly constrained 400ppm mineralisation

model (averaging 2.75t/m³). Table 14.4.3_1 shows the raw density statistics for mineralised

material (>100ppm U3O8), with Figures 14.4.3_1 and 14.4.3_2 showing histogram plots of the

density data. A density of 2.72t/m³ was used for fresh material and 2.0t/m³ was used for the

weathered zone.

Further density measurements are recommended for mineralised core intervals. There is

scope that the density applied to the model is conservative as the limestone unit itself has an

average density of 2.86t/m³ (from 7 readings).



Table 14.4.2_1

Berlin Project

Summary Mineralised Zone Composite Statistics

U3O8 (ppm)

P2O5 (%)

V2O5 (ppm)

Y2O3 (ppm)

Mo (ppm)

Ni (ppm)

Ag (ppm)

Re (ppm)

Nd (ppm)

Count 313 225 239 239 239 239 239 239 220

Minimum 32.15 0.8 230 34 18 39 0.36 0.1 10.8

Maximum 3290 21.5 11671 13740 1610 9070 25.0 37.8 316

Mean 1082 8.9 4623 483 593 2285 3.2 6.0 100

Median 1026 8.29 4490 498 587 2019 2.8 5.4 97.5

Standard Deviation 609 4.2 2181 273 304 1543 2.4 4.1 62.3

Variance 370,635 17.8 4,755,828 74,645 92,375 2,379,839 5.8 16.6 3,889

Standard Error 1.95 0.02 9.12 1.14 1.27 6.455 0.01 0.02 0.28

Coefficient of Variation 0.56 0.48 0.47 0.56 0.51 0.675 0.75 0.67 0.62

Correlation to Uranium

Pearson C.C. 0.85 0.81 0.88 0.63 0.64 0.5 0.69 0.77

Pearson Best Fit y = 55.4x + 2.85 y = 2.78x + 1,590.46 y = 0.38x + 67.84 y = 0.30x + 265.90 y = 1.56x + 581.59 y = 0.002x + 1.16 y = 0.004x + 1.16 y = 0.07x + 19.88

Spearman C.C. 0.88 0.86 0.89 0.72 0.75 0.72 0.83 0.78

Spearman Best Fit y = 0.88x + 13.90 y = 0.86x + 17.09 y = 0.89x + 13.27 y = 0.72x + 33.92 y = 0.75x + 29.91 y = 0.72x + 33.33 y = 0.83x + 19.95 y = 0.78x + 24.11



Figure 14.4.2_1 Multi Element Histogram Plots – Mineralised Zone

0

1

2

3

4

5

6

0 1000 2000 3000

Freq

uen

cy (

%)

resc_u3o8 (g/t)

Histogram Plot(Mineralised Zone)

0

1

2

3

4

5

6

7

8

9

10

11

-1 0 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22

Freq

uen

cy (

%)

P2O5_pct


0

1

2

3

4

1000 2000 3000 4000 5000 6000 7000 8000 9000 10000 11000

Freq

uen

cy (

%)

V2O5_ppm


0

1

2

3

4

5

6

0 100 200 300 400 500 600 700 800 900 1000 1100 1200 1300

Freq

uen

cy (

%)

Y2O3_ppm


0

1

2

3

4

5

0 100 200 300 400 500 600 700 800 900 1000 1100 1200 1300 1400 1500 1600

Freq

uen

cy (

%)

Mo_ppm


0

1

2

3

4

5

6

0 1000 2000 3000 4000 5000 6000 7000 8000 9000

Freq

uen

cy (

%)

Ni_ppm


0

2

4

6

8

10

12

14

16

18

20

22

24

26

28

-1 0 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25

Freq

uen

cy (

%)

Ag_ppm


0

2

4

6

8

10

12

14

16

18

0 10 20 30

Fre

qu

en

cy

(%

)

Re_ppm


0

1

2

3

4

100 200 300

Fre

qu

en

cy

(%

)

Nd_ppm




Figure 14.4.2_2 Scatter Plots for U3O8 (ppm) Compared to Other Elements

P2O5 (ppm)

V2O5 (ppm)

Y2O3 (ppm)

Mo (ppm)

0

20000

40000

60000

80000

100000

120000

140000

160000

180000

200000

220000

0 1000 2000 3000 400

P2O

5 (

ppm

)

resc_u3o8 (ppm)

P.CC= 0.85 S.CC= 0.88 Ref. Line y = 55.40x + 28,483.10

0

1000

2000

3000

4000

5000

6000

7000

8000

9000

10000

11000

12000

0 1000 2000 3000 400

V2O

5 (

ppm

)

resc_u3o8 (ppm)

P CC= 0 81 S CC= 0 86 Ref Line y = 2 78x + 1 590 46

0

1000

2000

3000

4000

0 1000 2000 3000 4000

Y2O

3 (

ppm

)

resc_u3o8 (ppm)

P.CC= 0.88 S.CC= 0.89 Ref. Line y = 0.38x + 67.84

0

1000

2000

3000

4000

0 1000 2000 3000 4000

Mo

(ppm

)

resc_u3o8 (ppm)




Figure 14.4.2_3 Scatter Plots for U3O8 (ppm) Compared to Other Elements

Ni (ppm)

Ag (ppm)

Nd (ppm)

Re (ppm)

0

1000

2000

3000

4000

5000

6000

7000

8000

9000

10000

0 1000 2000 3000 4000

Ni (

ppm

)

resc_u3o8 (ppm)


0

10

20

30

40

0 1000 2000 3000 4000

Ag

(ppm

)

resc_u3o8 (ppm)


0

100

200

300

400

0 1000 2000 3000 4000

Nd

(ppm

)

resc_u3o8 (ppm)


0

10

20

30

40

0 1000 2000 3000 4000

Re

(ppm

)

resc_u3o8 (ppm)




Table 14.4.3_1

Berlin Project

Bulk Density Data Summary Statistics (t/m³) (Diameter Method)

Region Count Average U3O8

(ppm) Min

Density Max

Density Average Density

Mineralisation 27 263 1.67 3.08 2.72 Limestone (LMS) 32 N/A 2.1 3.27 2.86 Shale (SHL) 15 N/A 1.73 3.0 2.39 Carbonaceous Shale (SHC) 32 N/A 1.21 4.77 2.76 Sandstone (SND) 65 N/A 1.67 3.17 2.63

15.4.4 127B127BVariography

In this document, the term ‘variogram’ is used as a generic word to designate the function

characterising the variability of variables versus the distance between two samples. The Isatis

geostatistical software was used to analyse the Berlin variography. Both traditional semi-

variogram and correlograms were used to analyse the spatial variability of the composites for

the mineralised zones.

Due to the complex nature of the folding, and the overall broad drillhole spacing, it was

decided to model an omnidirectional variogram for the mineralised layer. The resulting U3O8

variogram is shown in Figure 14.4.4_1.

The resulting variogram showed an overall good structure and generally long ranges, however

due to the drillhole spacing there was poor local-scale information. The variogram model

parameters used in the estimate is shown in Table 14.4.4_1.



Figure 14.4.3_1 Berlin Density Histograms (Diameter Method)

0

10

20

30

40

1.4 1.5 1.6 1.7 1.8 1.9 2.0 2.1 2.2 2.3 2.4 2.5 2.6 2.7 2.8 2.9 3.0 3.1 3.2 3.3

Freq

uen

cy (

%)

diameter (t/m3)

Histogram Plot(Mineralisation)

0

10

20

30

40

50

1.8 1.9 2.0 2.1 2.2 2.3 2.4 2.5 2.6 2.7 2.8 2.9 3.0 3.1 3.2 3.3 3.4 3.5

Freq

uen

cy (

%)

diameter (t/m3)

Histogram Plot(Limestone)

0

10

20

30

1.4 1.5 1.6 1.7 1.8 1.9 2.0 2.1 2.2 2.3 2.4 2.5 2.6 2.7 2.8 2.9 3.0 3.1 3.2 3.3

Freq

uen

cy (

%)

diameter (t/m3)

Histogram Plot(Shale)



Figure 14.4.3_2 Berlin Density Histograms (Diameter Method)

0

10

20

30

40

50

1 2 3 4

Freq

uen

cy (

%)

diameter (t/m3)

Histogram Plot(Carbonaceous Shale)

0

10

20

30

40

1.4 1.5 1.6 1.7 1.8 1.9 2.0 2.1 2.2 2.3 2.4 2.5 2.6 2.7 2.8 2.9 3.0 3.1 3.2 3.3

Freq

uen

cy (

%)

diameter (t/m3)

Histogram Plot(Sandstone)



Figure 14.4.4_1 Omnidirectional U3O8 Variography

Table 14.4.4_1

Berlin Project

Omnidirectional Variogram Model Parameters for 0.8m U3O8 (ppm) Composites

Region Nugget C1 Major Range

Semi-Major Range

Minor Range

C2 Major Range

Semi-Major Range

Minor Range

Mineralisation 10% 65% 60 60 60 25% 180 180 180

15.5 80B80BBlock Model Construction

A block model was created using SurpacTM mining software with a parent cell size of 4m

(Easting) by 50m (Northing) by 40m (RL) which was sub-blocked down to 0.5m (Easting) by

12.5m (Northing) by 2.5m (RL). A rotation of 330° was applied to the block model. The block

model parameters are summarised below in Table 14.5_1 and the main variables coded into

the block model are shown below in Table 14.5_2.



Table 14.5_1

Berlin Project

Block Model Parameters

Easting (X) Northing (Y) RL (Z)

Origin 503500 616000 280

Extent 2000 4000 800

Block size (m) 4 50 40

Sub Block size (m) 0.5 12.5 2.5

Rotation 330° 0° -30º

Table 14.5_2

Berlin Project

Block Model Main Variables

Attribute Name Type Decimals Default Description

estflag Integer 0 Estimation pass

category Integer 0 Resource Classification; 1 = Measured, 2 = Indicated, 3 – Inferred, 0 = Unclassified

U3O8_cut_ok real 3 -99 U3O8 (ppm), OK estimate

P2O5_id2 real 3 -99 P2O5 (ppm) ID2 estimate

V2O5_id2 real 3 -99 V2O5 (ppm), ID2 estimate

Y2O3_id2 real 3 -99 Y2O3 (ppm), ID2 estimate

Mo_id2 real 3 -99 Mo (ppm), ID2 estimate

Ni_id2 real 3 -99 Ni (ppm), ID2 estimate

Au_id2 real 3 -99 Au (ppm), ID2 estimate

Re_id2 real 3 -99 Re (ppm), ID2 estimate

Nd_id2 real 3 -99 Nd (ppm), ID2 estimate

density real 3 2.72 Insitu Dry Bulk Density;2.0 for weathered material; 0 for Air

Zone Integer -99 Mineralisation = 1

15.6 81B81BGrade Estimation Parameters

15.6.1 128B128BU3O8 Grade Estimate

U3O8 (ppm) grade was estimated into the block model using Ordinary Block Kriging (‘OK’).

Sample neighbourhood testing was conducted using Surpac to determine an appropriate

search strategy for the OK estimation. The neighbourhood testing included investigations into

the minimum and maximum number of samples used for the estimation, block discrimination,

negative kriging weights, the slope of the regression and resulting kriging variance. The

variogram parameters used for the estimation were based upon the variography discussed in

Section 14.4.4.

A three pass search strategy was used to estimate the U3O8 grade data into the mineralised

zone. The sample search parameters are outlined in Table 14.6.1_1. A sample search

anisotropy was introduced to reflect the main trend of the synclinal structure that hosts the

mineralisation.

The resulting model was reviewed statistically, visually, and against northing swath-plots and

was considered to represent the informing data appropriately. The block model volume was

also checked against the mineralisation DTM to ensure volume consistency.



Table 14.6.1_1

Berlin Project

U3O8 Sample Search Parameters – Ordinary Kriging

Zones Pass

Search Orientation Search Radii Number of Samples

Major Semi-Major

Minor Major Axis (m)

Semi-Major

Axis (m)

Minor Axis (m)

Min Max Max / Hole

Mineralisation

1

0 330° 00° 00°

100 67 67 8 16 4

2 200 133 133 8 16 4

3 400 267 267 3 12 4

Discretisation of 5 x 5 x 3 used

15.6.2 129B129BMulti-Element Data

As there was generally less multi-element data available for the estimate, primarily due to the

intervals which relied upon radiometric grade data, it was decided to estimate the remaining

multi-element grade data using inverse distance squared to the power of 2 (‘ID2’).

Table 14.6.2_1 summarises the search parameters used for the ID2 estimate.

Table 14.6.2_1

Berlin Project

Multi-Element Sample Search Parameters – ID2

Zones Pass

Search Orientation Search Radii Number of Samples

Major Semi-Major

Minor Major Axis (m)

Semi-Major

Axis (m)

Minor Axis (m)

Min Max Max / Hole

Mineralisation 1

0 330° 00° 00° 200 67 67 4 12 3

2 500 335 335 2 12 3

Discretisation of 3 x 3 x 3 used

15.7 82B82BBulk Density

The bulk density values used for the resource model were based upon the data analysed in

Section 14.4.1. A value of 2.72t/m³ for fresh material and 2.0t/m³ for weathered material was

used within the modelled mineralised zone.

15.8 83B83BResource Reporting and Classification

15.8.1 130B130BIntroduction

The resource estimate for the Berlin uranium deposit was categorised in accordance with the

criteria laid out in the Canadian National Instrument 43-101 (‘43-101’) and the JORC Code. A

combination of Indicated and Inferred Resources has been identified using definitive criteria

determined during the validation of the grade estimates, with detailed consideration of the 43-

101 categorisation guidelines.



15.8.2 131B131BCriteria for Resource Categorisation

The resource for the Berlin Project has been classified as a combination of Indicated and

Inferred Mineral Resources based on the confidence level of the key criteria that were

considered during resource classification as presented in Table 14.8.2_1. Figure 14.8.2_1

illustrates the classification applied to the resource.

Table 14.8.2_1

Berlin Project - Colombia

Confidence Levels of Key Categorisation Criteria

Items Discussion Confidence

Drilling Techniques Diamond – industry standard approach. Some regions of poor recovery. High

Logging Standard nomenclature applied with recording and apparent high quality. High

Drill Sample Recovery Acceptable recoveries determined for the majority of the drilling. Moderate

Sub-sampling Techniques and Sample Preparation

Industry standard for diamond drilling. Moderate to High

Quality of Assay Data

Good internal laboratory and external quality control data available for the majority of the chemical assaying.

Moderate to High

More bulk density data is required. Moderate

Verification of Sampling and Assaying

QAQC analysis is within industry acceptable standards. Moderate to High

Location of Sampling Points The bulk of drill collars surveyed by DGPS and most drillholes have been downhole surveyed.

High

Data Density and Distribution Nominal 50m by 100m to 200m by 300m spacing. Moderate to Low

Audits or Reviews Coffey Mining has reviewed the site drilling and sampling procedures. High

Database Integrity No material errors identified. High

Geological Interpretation The interpreted lithological and mineralisation boundaries are considered robust and of good confidence.

Moderate

Estimation and Modelling Techniques

Estimates based on detailed statistical and geostatistical analysis. Moderate to High

Cutoff Grades 400ppm mineralisation targeted. Moderate

Mining Factors or Assumptions

Whole block estimates for all mineralised regions completed. No mining has been undertaken in the deposit. Open pit and underground mining is conceptualised for this project.

Moderate

Indicated Resources

Blocks were classified as Indicated in a small region which had well established geological

continuity and a nominal data spacing of 60m x 130m and which had a consistent grade profile.

Inferred Resources

Blocks not classified as Indicated and which had drilling within ~200m were classified as

Inferred.

Unclassified Estimate

Portions of the estimate were not classified in areas poorly defined by very broad spaced

drilling. Separate 3D shapes were used to define these regions. As this portion of the model

is not classified, the corresponding estimate is not suitable for public reporting and is not

tabulated as part of the resource.



Figure 14.8.2_1 Drillhole Locations, Mineralisation Model and Classification

Plan View

Oblique View



15.8.3 132B132BGrade Tonnage Reporting

The reported resource for the Berlin uranium deposit using various cutoff grades is

summarised below (Table 14.8.3_1).

Table 14.8.3_1

Berlin Project, Colombia –

January 17 Resource Estimate, 2012

Reported above various U3O8 lower cutoff grades using a bulk density of 2.72t/m³ for fresh material and 2.0t/m³ for weathered material

Ordinary Kriged Estimate for U3O8, multi-element data estimated using Inverse Distance to the power of 2, using 0.8m assay composite data

Parent Block of 50m(Y) x 4m (X) by 40m (Z) Preferred Reporting Cutoff – 0.04% U3O8

Lower Cutoff

(% U3O8) Mt

U3O8 %

Contained P2O5 (%)

V2O5 (%)

Y2O3 (ppm)

Mo (ppm)

Ni (%)

Ag (ppm)

Re (ppm)

Nd (ppm) U3O8

(Mkg) U3O8 (MLb)

Inferred

0.04 8.1 0.11 9.0 19.9 9.4 0.5 500 620 0.2 3.4 6.8 100

0.05 8.0 0.11 9.0 19.7 9.4 0.5 500 620 0.2 3.3 6.8 100

0.06 8.0 0.11 8.9 19.7 9.4 0.5 500 620 0.2 3.3 6.8 100

0.07 7.9 0.11 8.9 19.5 9.5 0.5 510 620 0.2 3.3 6.8 100

0.08 7.7 0.11 8.7 19.2 9.5 0.5 510 630 0.2 3.3 6.9 100

0.09 6.8 0.12 7.9 17.5 9.7 0.5 520 630 0.2 3.4 7.0 110

0.1 5.6 0.12 6.8 15.0 10.0 0.5 540 650 0.2 3.5 7.2 110

Indicated

0.04 0.6 0.11 0.7 1.5 8.4 0.4 460 570 0.2 2.8 6.1 110

0.05 0.6 0.11 0.7 1.5 8.4 0.4 460 570 0.2 2.8 6.1 110

0.06 0.6 0.11 0.7 1.5 8.4 0.4 460 570 0.2 2.8 6.1 110

0.07 0.6 0.11 0.7 1.5 8.4 0.4 460 570 0.2 2.8 6.1 110

0.08 0.6 0.11 0.7 1.5 8.4 0.4 460 570 0.2 2.8 6.1 110

0.09 0.6 0.11 0.7 1.5 8.4 0.4 460 580 0.2 2.9 6.1 110

0.1 0.5 0.11 0.6 1.2 8.6 0.4 480 590 0.2 2.9 6.3 110

Coffey Mining is unaware of any mining, metallurgical, infrastructure or other relevant factors

which may materially affect the resource. The availability of suitable water and power supplies

may be key factors in any future mining studies. Security factors may influence future operational

decisions, however the author understands that regionally the security situation is improving.

15.9 84B84BConclusions

The overall geometry and trend of the mineralisation is well defined but further infill

drilling is required to adequately define the areas of complex folding and structure.

Further investigations are required to determine if faulting is affecting the mineralisation,

particularly for portions considered suitable for underground mining.

Further density data is required within the mineralised zone.



16 14B14BMINERAL RESERVE ESTIMATES

No reserves have been stated for this project.



17 15B15BMINING METHODS

No mining studies have been undertaken for this project. It is conceptually envisaged that

mining of this deposit could be possible using open pit methods for near-surface mineralisation,

and underground mechanised mining methods for the deeper portions of the deposit.



18 16B16BRECOVERY METHODS

Not Applicable.



19 17B17BPROJECT INFRASTRUCTURE

No mining studies have been undertaken for the project. Road access is good with the

project area lying 59km from the paved Medellin - Bogota highway. A secondary paved road

leads 50km west from the Bogota-Medellin to the village of Norcasia. The road continues,

unpaved, 9km beyond to Berlin village.

The 395MW La Miel hydroelectric dam is located approximately 12km from the central part of

the project area.

Future studies will need to assess the level of infrastructure requirements for the project.



20 18B18BMARKET STUDIES AND CONTRACTS

U3O8 Corp. has not currently undertaken any market studies for the project.



21 19B19BENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT

As the Berlin Project is currently in the exploration stage, there have been no specific

requirements for environmental studies and social or community-related plans; however all

exploration work must be done in accordance with environmental guidelines (see Environmental

License under Section 4.1.6). Nonetheless, U3O8 Corp. has been pro-active in undertaking

social and environmental initiatives. These initiatives include providing local employment,

education campaigns and information workshops to discuss exploration plans with local leaders,

landowners and residents, establishing agricultural and small-scale fish producing projects as

alternate sources of food and income for local communities, supporting local school and

community activities as well as minimizing environmental impact by limiting the number of roads

through the use of a cable system to transport equipment and personnel around the project,

practicing health and safety measures such as protective gear and radiation monitoring for all

employees. Biodegradable lubricants and additives were used in the drilling. Drill water passes

through settling tanks prior to being re-used or discharged.

Nurseries have been established for the propagation of indigenous flora that is used in the

reclamation of drill platforms and other areas that require revegetation, such as the drainage

from which the Berlin village draws its potable water.



22 20B20BCAPITAL AND OPERATING COSTS

Not applicable



23 21B21BECONOMIC ANALYSIS

Not applicable



24 22B22BADJACENT PROPERTIES

Not applicable.



25 23B23BOTHER RELEVANT DATA AND INFORMATION

To the best of the author’s knowledge, all relevant information has been discussed in the

document. The success of the project will also depend, in part, on the local security situation

of the region.



26 24B24BINTERPRETATION AND CONCLUSIONS

The Berlin Project comprises stratabound uranium and associated mineralisation at the contact

between a sandstone unit and overlying black, carbonaceous mudstone unit of a Lower

Cretaceous sequence. Mineralisation is predominantly contained within a phosphorous-rich

limestone layer within a north-north-east to north trending syncline. Typically, little significant

mineralisation occurs in the surrounding sediments. The mineralised layer ranges between

0.6m and 7m thick and averages 3m.

The uranium mineralisation occurs as secondary minerals from acidic, hydrocarbon-bearing fluid

dissolving rock and feldspar grains creating secondary permeability in the clastic carbonate

host-rocks, while simultaneously precipitating hydrocarbon and phosphate thereby roughly

maintaining the volume of the original rock.

The overall mineralisation geometry has been identified through diamond drilling and surface

trenching but still requires further delineation especially in areas of complex folding and to

determine if faulting is affecting the mineralisation zone.

U3O8 Corp. has completed a successful exploration program that has delineated a maiden

Resource for the project area.

A maiden Resource has been calculated for the Berlin Project (using a U3O8 cutoff of

400ppm U3O8.) of 8.1Mt at 0.11% U3O8 of Inferred material and 0.6Mt @ 0.11% U3O8 of

Indicated material. The success of the project will also depend, in part, on the local security

situation of the region.

Coffey has reviewed the drilling and QAQC data and considers it to be of a high standard for

use in the resource estimate.

Metallurgical testwork has investigated a number of means of extracting uranium, phosphate

and associated metals of potential economic interest. Two principal process routes have

potential to be applied at Berlin. The first, which has obtained outstanding recoveries for a

wide range of elements of potential economic interest, including uranium, at Berlin, includes

ferric leach of milled ore followed by a dilute acid leach. The second method that shows

promise for application at Berlin uses a pre-flotation step to generate an organic carbon – rich

concentrate that contains the majority of the uranium. An acid leach on the carbon-rich

concentrate appears to be effective in recovering uranium. Additional flotation tests are

underway to optimise the pre-flotation step and to define a flotation process for the resulting

carbon-poor material that will separate calcite and apatite so that apatite-rich material can be

subjected to leaching for metal recovery and/or the production of a phosphate product.



27 25B25BRECOMMENDATIONS

The following recommendations are made to advance the Berlin Project:

Additional drilling information is required to adequately define the areas of complex folding

and structure; determine if faulting is affecting the mineralisation, particularly for portions

considered suitable for underground mining; determine the northern extent of the

mineralisation. Further work is required to appropriately drill these holes, and any future

drilling programs will likely be guided by corporate objectives for the project. Accordingly,

the budget given below reflects a nominal program of 15 to 20 drillholes.

Further bulk density data is required to better quantify the density characteristics from

within the mineralised zone.

While chemical assaying is the preferred method for the Berlin Project, Coffey suggests

further analysis of radiometric eU3O8 data is required to improve the understanding of the

radiometric disequilibrium. A greater number of intervals with both chemical assays and

radiometric eU3O8 data are required to achieve this.

U3O8 Corp. has used good high grade standard reference materials, but Coffey

recommends using lower grade uranium standards of between 500ppm and 1,200ppm U3O8

to assess the precision of the lower grade material as well.

Further work is required to understand and model the depth of the weathering profile, and

assess the metallurgical properties of the weathered material.

Continue with testwork designed to establish optimal conditions for ferric iron leach of

whole ore. This work should focus on defining the effect of lower ferric iron concentrations

in solution as well as higher pulp densities as a means of reducing reagent consumption.

Additional flotation testwork is required to build on the success of certain aspects of

flotation, including the generation of cleaner sulphide and carbon pre-concentrates.

Work should continue on the separation of calcite and apatite through flotation in order to

provide a means of further reducing acid and/or ferric iron consumption.

A breakdown of projected costs to achieve the above recommendations is summarised in

Table 26_1.

Table 26_1

Costing for the Berlin Project Recommendations

Item Estimated Cost

(C$)

Additional Exploration and infill drilling and assaying $2.3M

Further Metallurgical testwork $0.5M

QAQC standards $0.01M

Additional density $0.02M

Additional resource Studies $0.1M

Total $2.93M



28 26B26BREFERENCES

Alvarez, J.A., 1979. Geología de la Cordillera Central y el Occidente colombiano y petroquímica de los intrusivos granitoides Meso-Cenozoicos. Unpubl. PhD thesis, Univ. Chile.

ANSTO, 2011. Mineralogy of the Berlin compsite. Unpubl. Technical Memorandum prepared for U3O8 Corp.., 39pp.

Barrero, D., and J. Vesga, 1976. Mapa geológico de cuadrángulo K-9, Armero y parte sur del J-9, La Dorada (1:1000 000): Ingeominas, Bogotá.

Bouma, A.H. 1962. Sedimentology of Some Flysch Deposits. Amsterdam: Elsevier Pub. Co., 168pp.

Botero, G. 1963. Contribucion al conocimiento de la geologia de la zona central de Antioquia. Anuls of the Faculty of Mines, Medellin: Vol 57, 101pp.

Bürgl, H., and L. Radelli, 1962. Nuevas localidades fosilíferas en la Cordillera Central de Colombia: Geología Colombiana, v.3, p.133-138.

Castaño, R., 1981. Calcul proviso ire des reserves geologiques de Berlin, sur la base des resulltants des sondages, unpubl. Minatome report in French, 15pp.

Cediel, F., Shaw, R.P. and Cáceres, C., 2003. Tectonic Assembly of the Northern Andean Block. In: C. Bartolini, R.T. Buffler and J. Blickwede, eds., The Circum-Gulf of Mexico and the Caribbean: Hydrocarbon habitats, basin formation and plate tectonics: AAPG Memoir 79, p.815-848.

Caceres, A., in preparation. Genesis of the sediment-hosted uraniferous phosphate deposit in the Berlin Project, Central Cordillera, Colombia and its implications for exploration. Unpubl. M.Sc. thesis, Queen's University, ON, Canada.

Cáceres, C., F. Etayo-Serna, and F. Cediel, 2003. Map 16, in C. Cáceres, F. Cediel, and F. Etayo, eds., Maps of sedimentary facies distribution and tectonic setting of Colombia through the Proterozoic and Phanerozoic: Bogotá, Ingeominas, p.41.

Dueñas, H., and E. Castro, 1981. Asociación palinológica de la Formación Mesa en la región Salán - Tolima, Colombia: Geología Norandina, Bogotá, v.3, p.27-36.

Etayo Serna, F., D. Barrero, H. Lozano, A. Espinosa, H. González, A. Orrego, I. Ballesteros, H. Forero, C. Ramírez, F. Zambrano-Ortiz, H. Duque-Caro, R. Vargas, A. Núñez, J. Álvarez, C. Ropaín, E. Cardozo, N. Galvis, L. Sarmiento, J. P. Alberts, J.E. Case, D.A. Singer, R. W. Bowen, B.R. Berger, D.P. Cox, and C.A. Hodges, 1986. Mapa de terrenos geológicos de Colombia: Publicaciones Especiales del Ingeominas, v.14, p.1-135.

Etayo-Serna, F., 1985. Documentación paleontológica del infracretácico de San Félix y Valle Alto, Cordillera Central, in F. E. Serna, and F. L. Montaño, eds., Proyecto Cretácico, v. 16: Bogotá, Publicaciones Geológicas Especiales del Ingeominas, p.XXV-1 XXV-7.

Etayo-Serna, F., C. Cáceres, and F. Cediel, 2003. Map 5 - Map 8, in C. Cáceres, F. Cediel, and F. Etayo, eds., Maps of sedimentary facies distribution and tectonic setting of Colombia through the Proterozoic and Phanerozoic: Bogotá, Ingeominas, p.18-25.



Feininger, T., D. Barrero, and N. Castro, 1972. Geología de parte de los departamentos de Antioquia y Caldas (Sub-zona II-B): Boletín Geológico Ingeominas, v.20, p.1-173.

Fetter, A.H, Santos, T.J.S., Van Schmus, W.R., Hackspacher, P.C., Brito Neves, B.B., Arthaud, M.H. Nogueira Neto, J.A. & Wernick, E. 2003. Evidence for Neoproterozoic continental arc magmatism in the Santa Quiteria Batholith of Ceara State, NW Borborema Province, NE Brazil: implications for the assembly of west Gondwana. Gondwana Research, 6: 265-273.

Gharabaghi, M., Irannajad. M. and Noaparast, M., 2010. A review of the beneficiation of calcareous phosphate ores using organic acid leaching. Hydrometallurgy 103: 96–107.

González, H., 1980. Geología de las planchas 167 (Sonsón) y 187 (Salamina): Boletín Geológico Ingeominas, v.23, p.1-174.

Guerrero, J., 1993. Magnetostratigraphy of the upper part of the Honda Group and Neiva Formation: Miocene uplift of the Colombian Andes, Duke University.

Hettner, A., 1982. Die Kordilleren Von, Bogotá. Peterm. Mitt. Erg. Bd. 22, Heft 104, p.1-131, Ingeominas, Bogotá.

Maya, M., 1992. Catálogo de dataciones isotópicas en Colombia: Boletín Geológico Ingeominas, v.32, p.127-188.

Maya, M., and H. González, 1995. Unidades litodémicas en la Cordillera Central de Colombia: Boletín Geológico Ingeominas, v.35, p 43-57.

Mining Journal, 2009. Sweden - Supplement to the Mining Journal, London. 12pp.

Moreno-Sánchez, M., A. Gómez Cruz, and H. Castillo González, 2008. Ocurrencias de fósiles paleozoicos al este de la parte norte de la Cordillera Central y discusión sobre su significado geológico: Boletín de Ciencias de la Tierra, v.22, p.39-47.

Muñoz, C.A., 1983. Determinación del potencial uranífero de la Alaskita de Samaná, departamento de Caldas - Colombia: Trabajo de grado, Universidad Nacional de Colombia, Bogotá D. C., 75 p.

Naranjo, J.L., 1983. Investigación del potencial uranífero en los shales negros del Sinclinal de Berlín, departamento de Caldas: Trabajo de grado, Universidad Nacional de Colombia, Bogotá D. C., 114 p.

Nelson, H.W., 1957. Contribution to the geology of the Central and Western Cordillera of Colombia in the sector between Ibagué and Cali, Eduard Ijdo, Leiden.

Núñez, T.A., H. González, and E. Linares, 1979. Nuevas edades K/Ar de los esquistos verdes del Grupo Cajamarca: Publicaciones Especiales de Geología, v.23, p.18.

Renaud, J., 2010a. A petrographic and microprobe investigation of samples from trenches from the southern part of the Berlin syncline, Caldas Province, Colombia. Unpubl. Internal Report for U3O8 Corp.., 53pp.

Renaud, J., 2010b. A petrographic and microprobe investigation of core samples from drillholes DDB5 and DDB7 from the southern part of the Berlin Syncline, Caldas Province, Colombia. Unpubl. Internal Report for U3O8 Corp.., 36pp.



Renaud, J., 2010c. A petrographic and microprobe investigation of samples from drillholes DDB1 and DDB3, from the southern part of the Berlin Syncline, Caldas Province, Colombia. Unpubl. Internal Report for U3O8 Corp.., 32pp.

Restrepo, J.J., J. Toussaint, H. González, U. Cordani, K. Kawashita, E. Linares, and C. Parica, 1991. Precisiones geocronológicas sobre el occidente colombiano: Simposio sobre Magmatismo Andino y su marco tectónico, v. Memorias (Tomo 1), p.1-22.

Rodriguez Dominguez, H. 2012. Legal Audit Report – Gaia Energy. Unpublished report by Norton Rose. February 2012.

Roussemet, R. & Houot, R. 1979. Valorisation d’un phosphate uranifere. Unpublished Report, Minatome Colombiana Ltda and Instituto de Asuntos Nucleares, Special Report No 7, Bogota, Colombia.

Serna, G. 2011. Summary of the Environmental Activity of the Berlin Project. Unpublished Technical Report. Geologist Environmental Advisory.

SRK, 2006. Independent Technical Report on the KPS Ventures Uranium Exploration Properties in Colombia, Unpublished NI 43-101 Technical Report, SRK Consulting.

Thouret, J.C., 1989. Geomorfología y cronoestratigrafía del Macizo Volcánico Ruiz - Tolima (Cordillera Central Colombiana), in T. Van Der Hammen, S. Diaz-Piedrahita, V. J. Alvarez, eds., Estudios de Ecosistemas Tropoandinos, v3: Berlín, La Cordillera Central Colombiana – Transecto Parque Los Nevados (segunda parte), p.257-277.

Toussaint, J. and J.J. Restrepo, 1988. Son alóctonos los Andes Colombianos?: Revista I.C.N.E. Universidad Nacional Medellín, v.1, p.17-41.

Vakhrameev, V.A., 1991. Jurassic and Cretaceous floras and climates of the earth, Cambridge University Press, London, 340 p.

Appendix A QAQC Summary Plots

Berlin Project Uranium Standards Feb 2012

Appendix A – QAQC Summary Plots Page: 1

Standard: ST1000020 No of Analyses: 9Element: U Result Minimum: 1,760.00Units: ppm Maximum: 1,960.00Detection Limit: 1 Mean: 1,848.89Expected Value (EV): 1,910.20 Std Deviation: 66.07E.V. Range: 1,719.18 to 2,101.22 % in Tolerance 100.00 %

% Bias -3.21 %% RSD 3.57 %

17 00

18 00

19 00

20 00

21 00

22 00

15

-No

v-2

01

0

18

-No

v-2

01

0

18

-No

v-2

01

0

02

-De

c-201

0

02

-De

c-201

0

06

-De

c-201

0

16

-De

c-201

0

10

-Jan

-20

11

U_R

esu

lt (

ppm

)

DateReported

Standard Control Plot(Standard ST1000020 ME-MS61 Analysis )


-15 0

-10 0

-50

0

50

10 0

15

-No

v-2

01

0

18

-No

v-2

01

0

18

-No

v-2

01

0

02

-De

c-201

0

02

-De

c-201

0

06

-De

c-201

0

16

-De

c-201

0

10

-Jan-2

011C

um

ula

tiv

e S

um

of

U_

Resu

lt -

Mea

n (

pp

m)

DateReported

Cumulative Deviation from Assay Mean(Standard ST1000020 ME-MS61 Analysis )

U_Result Mean of Cumulative Sum of U_Result - Mean (ppm) = -24.44

-60 0-50 0-40 0-30 0-20 0-10 0

0

15

-No

v-2

01

0

18

-No

v-2

01

0

18

-No

v-2

01

0

02

-De

c-201

0

02

-De

c-201

0

06

-De

c-201

0

16

-De

c-201

0

10

-Jan

-20

11

Cu

mula

tiv

e S

um

of

U_

Resu

lt -

Ex

pect

ed

Va

lue (

p

DateReported

Cumulative Deviation from Expected Value(Standard ST1000020 ME-MS61 Analysis )

U_Result Mean of Cumulative Sum of U_Result - Expected Value (ppm) = -331.00

Berlin Project Standards Feb 2012(Standard ST1000020 ME-MS61 Analysis )

Printed: 03-Feb-2012 13:16:01 Data Imported: 02-Feb-2012 10:33:23 Page 1




% Bias -5.64 %% RSD 4.66 %

15 0016 0017 0018 0019 0020 0021 0022 00

02-F

eb

-2011

09-A

pr-2

011

01-Ju

l-201

1

18-A

ug

-201

1

U_R

esu

lt (

ppm

)

DateReported



-40 0

-30 0

-20 0

-10 0

0

10 00

2-F

eb

-201

1

09

-Ap

r-201

1

01

-Jul-2

011

18

-Au

g-2

011C

um

ula

tiv

e S

um

of

U_

Resu

lt -

Mea

n (

pp

m)

DateReported

Cumulative Deviation from Assay Mean(Standard ST1000020 ME-MS61U Analysis )


-50 00

-40 00

-30 00

-20 00

-10 00

0

02-F

eb

-201

1

09-A

pr-2

011

01-Ju

l-20

11

18-A

ug

-201

1

Cu

mula

tiv

e S

um

of

U_

Resu

lt -

Ex

pect

ed

Va

lue (

p

DateReported

Cumulative Deviation from Expected Value(Standard ST1000020 ME-MS61U Analysis )

U_Result Mean of Cumulative Sum of U_Result - Expected Value (ppm) = -2,627.04

Berlin Project Standards Feb 2012(Standard ST1000020 ME-MS61U Analysis )





% Bias -6.89 %% RSD 5.50 %

11 00

12 00

13 00

14 00

15 00

16 00

17

-Ma

y-2

01

1

01

-Jul-2

011

30

-Jul-2

011

26

-Sep

-201

1

U_R

esu

lt (

ppm

)

DateReported



-80 0

-60 0

-40 0

-20 0

0

20 01

7-M

ay-2

01

1

01

-Jul-2

011

30

-Jul-2

011

26

-Se

p-2

01

1Cu

mula

tive S

um

of

U_

Resu

lt -

Mea

n (

ppm

)

DateReported

Cumulative Deviation from Assay Mean(Standard ST1000045 ME-MS61U Analysis )


-50 00

-40 00

-30 00

-20 00

-10 00

0

17

-Ma

y-2

011

01

-Jul-2

01

1

30

-Jul-2

01

1

26

-Se

p-2

01

1

Cu

mula

tiv

e S

um

of

U_

Resu

lt -

Ex

pect

ed

Valu

e (

p

DateReported

Cumulative Deviation from Expected Value(Standard ST1000045 ME-MS61U Analysis )

U_Result Mean of Cumulative Sum of U_Result - Expected Value (ppm) = -2,292.51

Berlin Project Standards Feb 2012(Standard ST1000045 ME-MS61U Analysis )




Standard: AMIS-055 No of Analyses: 14Element: U Result Minimum: 3,200.00Units: ppm Maximum: 3,620.00Detection Limit: 1 Mean: 3,360.71Expected Value (EV): 3,206.00 Std Deviation: 112.15E.V. Range: 2,885.40 to 3,526.60 % in Tolerance 92.86 %

% Bias 4.83 %% RSD 3.34 %

28 00

30 00

32 00

34 00

36 00

38 00

20

-Jul-2

01

1

30

-Jul-2

01

1

04

-Au

g-2

01

1

13

-Au

g-2

01

1

18

-Au

g-2

01

1

19

-Au

g-2

01

1

26

-Au

g-2

01

1

05

-Se

p-2

01

1

19

-Se

p-2

01

1

19

-Se

p-2

01

1

19

-Se

p-2

01

1

01

-Oct-2

01

1

01

-Oct-2

01

1

U_R

esu

lt (

pp

m)

DateReported

Standard Control Plot(Standard AMIS-055 ME-MS61U Analysis )


-40 0-30 0-20 0-10 0

010 020 0

20

-Jul-2

01

1

30

-Jul-2

01

1

04

-Au

g-2

01

1

13

-Au

g-2

01

1

18

-Au

g-2

01

1

19

-Au

g-2

01

1

26

-Au

g-2

01

1

05

-Se

p-2

01

1

19

-Se

p-2

01

1

19

-Se

p-2

01

1

19

-Se

p-2

01

1

01

-Oct-2

01

1

01

-Oct-2

01

1Cu

mula

tiv

e S

um

of

U_

Resu

lt -

Mea

n (

pp

m)

DateReported

Cumulative Deviation from Assay Mean(Standard AMIS-055 ME-MS61U Analysis )


-50 00

50 010 0015 0020 0025 00

20

-Jul-2

01

1

30

-Jul-2

01

1

04

-Au

g-2

01

1

13

-Au

g-2

01

1

18

-Au

g-2

01

1

19

-Au

g-2

01

1

26

-Au

g-2

01

1

05

-Se

p-2

01

1

19

-Se

p-2

01

1

19

-Se

p-2

01

1

19

-Se

p-2

01

1

01

-Oct-2

01

1

01

-Oct-2

01

1

Cu

mula

tiv

e S

um

of

U_

Resu

lt -

Ex

pect

ed

Va

lue

(p

DateReported

Cumulative Deviation from Expected Value(Standard AMIS-055 ME-MS61U Analysis )

U_Result Mean of Cumulative Sum of U_Result - Expected Value (ppm) = 1,115.00

Berlin Project Standards Feb 2012(Standard AMIS-055 ME-MS61U Analysis )


Berlin Project Vanadium Standards Feb 2012


Standard: ST1000020 No of Analyses: 9Element: V Result Minimum: 3,100.00Units: ppm Maximum: 3,490.00Detection Limit: Mean: 3,274.44Expected Value (EV): 3,414.00 Std Deviation: 141.27E.V. Range: 3,072.60 to 3,755.40 % in Tolerance 100.00 %

% Bias -4.09 %% RSD 4.31 %

30 00

32 00

34 00

36 00

38 00

15

-No

v-2

01

0

18

-No

v-2

01

0

18

-No

v-2

01

0

02

-De

c-201

0

02

-De

c-201

0

06

-De

c-201

0

16

-De

c-201

0

10

-Jan

-20

11

V_R

esu

lt (

pp

m)

DateReported

Standard Control Plot(Standard ST1000020 V ME- MS61 Analysis)


-40 0

-30 0

-20 0

-10 0

0

10 0

15

-No

v-2

01

0

18

-No

v-2

01

0

18

-No

v-2

01

0

02

-De

c-201

0

02

-De

c-201

0

06

-De

c-201

0

16

-De

c-201

0

10

-Jan

-20

11C

um

ula

tiv

e S

um

of

V_

Resu

lt -

Mea

n (

pp

m)

DateReported

Cumulative Deviation from Assay Mean(Standard ST1000020 V ME- MS61 Analysis)

V_Result Mean of Cumulative Sum of V_Result - Mean (ppm) = -145.56

-14 00-12 00-10 00

-80 0-60 0-40 0-20 0

0

15

-No

v-2

01

0

18

-No

v-2

01

0

18

-No

v-2

01

0

02

-De

c-201

0

02

-De

c-201

0

06

-De

c-201

0

16

-De

c-201

0

10

-Jan

-20

11

Cu

mula

tiv

e S

um

of

V_

Resu

lt -

Ex

pect

ed

Va

lue

(p

DateReported

Cumulative Deviation from Expected Value(Standard ST1000020 V ME- MS61 Analysis)

V_Result Mean of Cumulative Sum of V_Result - Expected Value (ppm) = -843.33

Berlin Project Standards Feb 2012(Standard ST1000020 V ME- MS61 Analysis)





% Bias -6.16 %% RSD 2.99 %

30 00

32 00

34 00

36 00

38 00

02

-Fe

b-2

01

1

09

-Ap

r-201

1

01

-Jul-2

01

1

18

-Au

g-2

01

1

V_R

esu

lt (

pp

m)

DateReported



-60 0

-40 0

-20 0

0

20 0

40 00

2-F

eb

-201

1

09

-Ap

r-201

1

01

-Jul-2

01

1

18

-Au

g-2

01

1

Cu

mula

tiv

e S

um

of

V_

Resu

lt -

Mea

n (

pp

m)

DateReported

Cumulative Deviation from Assay Mean(Standard ST1000020 V ME- MS61U Analysis )

V_Result Mean of Cumulative Sum of V_Result - Mean (ppm) = 18.75

-10 000

-80 00

-60 00

-40 00

-20 00

0

02

-Fe

b-2

01

1

09

-Ap

r-201

1

01

-Jul-2

01

1

18

-Au

g-2

01

1

Cu

mula

tiv

e S

um

of

V_

Resu

lt -

Ex

pect

ed

Va

lue (

p

DateReported

Cumulative Deviation from Expected Value(Standard ST1000020 V ME- MS61U Analysis )

V_Result Mean of Cumulative Sum of V_Result - Expected Value (ppm) = -4,709.32

Berlin Project Standards Feb 2012(Standard ST1000020 V ME- MS61U Analysis )





% Bias -0.59 %% RSD 7.54 %

28 0030 0032 0034 0036 0038 0040 00

11-Ja

n-2

011

28-F

eb

-2011

19-A

pr-2

011

V_R

esu

lt (

ppm

)

DateReported

Standard Control Plot(Standard ST1000020 V ME- MS81Analysis )


-40 00

-30 00

-20 00

-10 00

0

10 00

11

-Jan

-2011

28

-Fe

b-2

01

1

19

-Ap

r-201

1

Cu

mula

tive S

um

of

V_

Resu

lt -

Mean

(p

pm

)

DateReported

Cumulative Deviation from Assay Mean(Standard ST1000020 V ME- MS81Analysis )

V_Result Mean of Cumulative Sum of V_Result - Mean (ppm) = -1,560.00

-40 00

-30 00

-20 00

-10 00

0

11

-Jan-2

011

28

-Fe

b-2

01

1

19

-Apr-2

01

1

Cu

mula

tive S

um

of

V_

Resu

lt -

Exp

ect

ed

Va

lue (

p

DateReported

Cumulative Deviation from Expected Value(Standard ST1000020 V ME- MS81Analysis )


U308Corp Berlin Project(Standard ST1000020 V ME- MS81Analysis )




Standard: ST1000045 No of Analyses: 41Element: V_Result Minimum: 3,020.00Units: ppm Maximum: 3,580.00Detection Limit: Mean: 3,348.54Expected Value (EV): 3,584.00 Std Deviation: 106.49E.V. Range: 3,225.60 to 3,942.40 % in Tolerance 90.24 %

% Bias -6.57 %% RSD 3.18 %

30 00

32 00

34 00

36 00

38 00

40 00

17

-Ma

y-2

01

1

01

-Jul-2

01

1

30

-Jul-2

01

1

26

-Se

p-2

01

1

V_R

esu

lt (

pp

m)

DateReported



-60 0-50 0-40 0-30 0-20 0-10 0

010 0

17

-Ma

y-2

01

1

01

-Jul-2

01

1

30

-Jul-2

01

1

26

-Se

p-2

01

1Cu

mula

tiv

e S

um

of

V_

Resu

lt -

Mea

n (

pp

m)

DateReported

Cumulative Deviation from Assay Mean(Standard ST1000045 V ME- MS61U Analysis )

V_Result Mean of Cumulative Sum of V_Result - Mean (ppm) = -187.07

-10 000

-80 00

-60 00

-40 00

-20 00

0

17

-Ma

y-2

01

1

01

-Jul-2

01

1

30

-Jul-2

01

1

26

-Se

p-2

01

1

Cu

mula

tiv

e S

um

of

V_

Resu

lt -

Ex

pect

ed

Va

lue

(p

DateReported

Cumulative Deviation from Expected Value(Standard ST1000045 V ME- MS61U Analysis )


Berlin Project Standards Feb 2012(Standard ST1000045 V ME- MS61U Analysis )




Standard: ST1000045 No of Analyses: 9Element: V_Result Minimum: 3,320.00Units: ppm Maximum: 4,030.00Detection Limit: Mean: 3,740.00Expected Value (EV): 3,584.00 Std Deviation: 233.71E.V. Range: 3,225.60 to 3,942.40 % in Tolerance 77.78 %

% Bias 4.35 %% RSD 6.25 %

32 00

34 00

36 00

38 00

40 00

42 00

23

-Fe

b-2

01

1

19

-Ap

r-201

1

19

-Ap

r-201

1

19

-Ap

r-201

1

21

-Ap

r-201

1

17

-Ma

y-2

01

1

26

-Ma

y-2

01

1

18

-Jul-2

01

1

V_R

esu

lt (

pp

m)

DateReported

Standard Control Plot(Standard ST1000045 V ME- MS81 Analysis )


-40 0

-20 0

0

20 0

40 0

60 0

23

-Fe

b-2

01

1

19

-Ap

r-201

1

19

-Ap

r-201

1

19

-Ap

r-201

1

21

-Ap

r-201

1

17

-Ma

y-2

01

1

26

-Ma

y-2

01

1

18

-Jul-2

01

1Cu

mula

tiv

e S

um

of

V_

Resu

lt -

Mea

n (

pp

m)

DateReported

Cumulative Deviation from Assay Mean(Standard ST1000045 V ME- MS81 Analysis )

V_Result Mean of Cumulative Sum of V_Result - Mean (ppm) = 28.89

-50 0

0

50 0

10 00

15 00

20 00

23

-Fe

b-2

01

1

19

-Ap

r-201

1

19

-Ap

r-201

1

19

-Ap

r-201

1

21

-Ap

r-201

1

17

-Ma

y-2

01

1

26

-Ma

y-2

01

1

18

-Jul-2

01

1

Cu

mula

tiv

e S

um

of

V_

Resu

lt -

Ex

pect

ed

Va

lue

(p

DateReported

Cumulative Deviation from Expected Value(Standard ST1000045 V ME- MS81 Analysis )

V_Result Mean of Cumulative Sum of V_Result - Expected Value (ppm) = 808.89

Berlin Project Standards Feb 2012(Standard ST1000045 V ME- MS81 Analysis )

Printed: 02-Mar-2012 15:21:23 Data Imported: 02-Feb-2012 12:54:00 Page 1

Berlin Project Molybdenum Standards Feb 2012


Standard: ST1000020 No of Analyses: 41Element: Mo Result Minimum: 47.90Units: ppm Maximum: 70.50Detection Limit: Mean: 57.85Expected Value (EV): 64.18 Std Deviation: 4.70E.V. Range: 57.76 to 70.60 % in Tolerance 56.10 %

% Bias -9.86 %% RSD 8.12 %

40

50

60

70

80

17

-Ma

y-2

01

1

01

-Jul-2

01

1

30

-Jul-2

01

1

26

-Se

p-2

01

1

Mo

_R

esu

lt (

pp

m)

DateReported

Standard Control Plot(Standard ST1000020 Mo ME-MS61U Analysis)

Mo_Result Expected Value = 64.18 EV Range (57.76 to 70.60) Mean of Mo_Result = 57.85

-50-40-30-20-10

010

17

-Ma

y-2

01

1

01

-Jul-2

01

1

30

-Jul-2

01

1

26

-Se

p-2

01

1Cu

mula

tiv

e S

um

of

Mo

_R

esu

lt -

Me

an (

pp

m)

DateReported

Cumulative Deviation from Assay Mean(Standard ST1000020 Mo ME-MS61U Analysis)

Mo_Result Mean of Cumulative Sum of Mo_Result - Mean (ppm) = -16.48

-30 0

-20 0

-10 0

0

17

-Ma

y-2

01

1

01

-Jul-2

01

1

30

-Jul-2

01

1

26

-Se

p-2

01

1

Cu

mula

tiv

e S

um

of

Mo

_R

esu

lt -

Exp

ect

ed

Va

lue

(

DateReported

Cumulative Deviation from Expected Value(Standard ST1000020 Mo ME-MS61U Analysis)

Mo_Result Mean of Cumulative Sum of Mo_Result - Expected Value (ppm) = -149.44

Berlin Project Standards Feb 2012(Standard ST1000020 Mo ME-MS61U Analysis)




Standard: ST1000020 No of Analyses: 44Element: Mo_Result Minimum: 52.90Units: ppm Maximum: 76.40Detection Limit: Mean: 65.59Expected Value (EV): 64.18 Std Deviation: 4.21E.V. Range: 57.76 to 70.60 % in Tolerance 84.09 %

% Bias 2.20 %% RSD 6.42 %

50

60

70

80

02

-Fe

b-2

01

1

09

-Ap

r-201

1

01

-Jul-2

01

1

18

-Au

g-2

01

1

Mo

_R

esu

lt (

pp

m)

DateReported

Standard Control Plot(Standard ST1000020 Mo ME-MS81 Analysis)


-40

-30

-20

-10

0

100

2-F

eb

-201

1

09

-Ap

r-201

1

01

-Jul-2

01

1

18

-Au

g-2

01

1Cu

mula

tiv

e S

um

of

Mo

_R

esu

lt -

Me

an (

pp

m)

DateReported

Cumulative Deviation from Assay Mean(Standard ST1000020 Mo ME-MS81 Analysis)


-20

0

20

40

60

80

02

-Fe

b-2

01

1

09

-Ap

r-201

1

01

-Jul-2

01

1

18

-Au

g-2

01

1

Cu

mula

tiv

e S

um

of

Mo

_R

esu

lt -

Exp

ect

ed

Va

lue

(

DateReported

Cumulative Deviation from Expected Value(Standard ST1000020 Mo ME-MS81 Analysis)

Mo_Result Mean of Cumulative Sum of Mo_Result - Expected Value (ppm) = 16.08

Berlin Project Standards Feb 2012(Standard ST1000020 Mo ME-MS81 Analysis)




Standard: ST1000020 No of Analyses: 9Element: Mo_Result Minimum: 61.20Units: ppm Maximum: 70.80Detection Limit: Mean: 66.71Expected Value (EV): 64.18 Std Deviation: 3.36E.V. Range: 57.76 to 70.60 % in Tolerance 88.89 %

% Bias 3.94 %% RSD 5.03 %

55

60

65

70

75

15

-No

v-2

01

0

18

-No

v-2

01

0

18

-No

v-2

01

0

02

-De

c-201

0

02

-De

c-201

0

06

-De

c-201

0

16

-De

c-201

0

10

-Jan

-20

11

Mo

_R

esu

lt (

pp

m)

DateReported

Standard Control Plot(Standard ST1000020 Mo ME-MS61 Analysis )


-8-6-4-2024

15

-No

v-2

01

0

18

-No

v-2

01

0

18

-No

v-2

01

0

02

-De

c-201

0

02

-De

c-201

0

06

-De

c-201

0

16

-De

c-201

0

10

-Jan

-20

11C

um

ula

tiv

e S

um

of

Mo

_R

esu

lt -

Me

an (

pp

m)

DateReported

Cumulative Deviation from Assay Mean(Standard ST1000020 Mo ME-MS61 Analysis )


5

10

15

20

25

15

-No

v-2

01

0

18

-No

v-2

01

0

18

-No

v-2

01

0

02

-De

c-201

0

02

-De

c-201

0

06

-De

c-201

0

16

-De

c-201

0

10

-Jan

-20

11

Cu

mula

tiv

e S

um

of

Mo

_R

esu

lt -

Exp

ect

ed

Va

lue

(

DateReported

Cumulative Deviation from Expected Value(Standard ST1000020 Mo ME-MS61 Analysis )

Mo_Result Mean of Cumulative Sum of Mo_Result - Expected Value (ppm) = 11.24

Berlin Project Standards Feb 2012(Standard ST1000020 Mo ME-MS61 Analysis )


Berlin Project Phosphorus Standards Feb 2012


Standard: ST1000020 No of Analyses: 33Element: P Result Minimum: 1.80Units: % Maximum: 1.94Detection Limit: Mean: 1.87Expected Value (EV): 1.76 Std Deviation: 0.03E.V. Range: 1.58 to 1.94 % in Tolerance 96.97 %

% Bias 6.23 %% RSD 1.76 %

1.5

1.6

1.7

1.8

1.9

11

-Jan

-20

11

28

-Fe

b-2

011

19

-Ap

r-201

1

P_

Re

sult

(%

)

DateReported

Standard Control Plot(Standard ST1000020 P ME-MS81 Analysis)


-0.25-0.20-0.15-0.10-0.050.000.05

11

-Jan

-20

11

28

-Fe

b-2

011

19

-Ap

r-201

1

Cu

mula

tiv

e S

um

of

P_R

esu

lt -

Me

an (

%)

DateReported

Cumulative Deviation from Assay Mean(Standard ST1000020 P ME-MS81 Analysis)

P_Result Mean of Cumulative Sum of P_Result - Mean (%) = -0.12

0

1

2

3

4

11

-Jan

-20

11

28

-Fe

b-2

011

19

-Ap

r-201

1

Cum

ula

tiv

e Su

m o

f P

_Re

sult

- E

xp

ect

ed V

alu

e (

DateReported

Cumulative Deviation from Expected Value(Standard ST1000020 P ME-MS81 Analysis)

P_Result Mean of Cumulative Sum of P_Result - Expected Value (%) = 1.74

Berlin Project Standards Feb 2012(Standard ST1000020 P ME-MS81 Analysis)

Printed: 03-Feb-2012 11:29:09 Data Edited: 02-Feb-2012 17:00:20 Page 1



Standard: ST1000020 No of Analyses: 37Element: P Result Minimum: 1.80Units: % Maximum: 1.94Detection Limit: Mean: 1.87Expected Value (EV): 1.76 Std Deviation: 0.03E.V. Range: 1.58 to 1.94 % in Tolerance 97.30 %

% Bias 6.36 %% RSD 1.75 %

1.5

1.6

1.7

1.8

1.9

27

-Jan

-20

11

12

-Ma

r-20

11

08

-Jun

-20

11

P_

Re

sult

(%

)

DateReported



-0.20

-0.15

-0.10

-0.05

0.00

0.05

27

-Jan

-20

11

12

-Ma

r-201

1

08

-Jun

-20

11

Cu

mula

tiv

e S

um

of

P_R

esu

lt -

Me

an

(%

)

DateReported

Cumulative Deviation from Assay Mean(Standard ST1000020 P ME-XRF12 Analysis )


0

1

2

3

4

5

27

-Jan

-20

11

12

-Ma

r-20

11

08

-Jun

-20

11

Cu

mula

tiv

e S

um

of

P_R

esu

lt -

Ex

pe

cte

d V

alu

e (

DateReported

Cumulative Deviation from Expected Value(Standard ST1000020 P ME-XRF12 Analysis )


Berlin Project Standards Feb 2012(Standard ST1000020 P ME-XRF12 Analysis )




Standard: ST1000045 No of Analyses: 9Element: P_Result Minimum: 2.45Units: % Maximum: 2.54Detection Limit: Mean: 2.49Expected Value (EV): 2.44 Std Deviation: 0.02E.V. Range: 2.20 to 2.68 % in Tolerance 100.00 %

% Bias 2.10 %% RSD 0.94 %

2.12.22.32.42.52.6

23

-Fe

b-2

01

1

19

-Apr-2

01

1

19

-Apr-2

01

1

19

-Apr-2

01

1

21

-Apr-2

01

1

17

-Ma

y-2

01

1

26

-Ma

y-2

01

1

18

-Jul-2

01

1

P_

Resu

lt (

%)

DateReported

Standard Control Plot(Standard ST1000045 P ME-MS81 Analysis )


-0.05-0.04-0.03-0.02-0.010.000.01

23

-Feb

-201

1

19

-Ap

r-201

1

19

-Ap

r-201

1

19

-Ap

r-201

1

21

-Ap

r-201

1

17

-Ma

y-2

01

1

26

-Ma

y-2

01

1

18

-Jul-2

01

1

Cu

mula

tiv

e S

um

of

P_R

esu

lt -

Mean

(%

)

DateReported

Cumulative Deviation from Assay Mean(Standard ST1000045 P ME-MS81 Analysis )


0.0

0.1

0.2

0.3

0.4

0.5

23

-Fe

b-2

01

1

19

-Ap

r-201

1

19

-Ap

r-201

1

19

-Ap

r-201

1

21

-Ap

r-201

1

17

-Ma

y-2

01

1

26

-Ma

y-2

01

1

18

-Jul-2

01

1

Cu

mula

tiv

e S

um

of

P_R

esu

lt -

Ex

pe

cte

d V

alu

e (

DateReported

Cumulative Deviation from Expected Value(Standard ST1000045 P ME-MS81 Analysis )


Berlin Project Standards Feb 2012(Standard ST1000045 P ME-MS81 Analysis )




Standard: ST1000045 No of Analyses: 20Element: P_Result Minimum: 2.45Units: % Maximum: 2.54Detection Limit: Mean: 2.49Expected Value (EV): 2.44 Std Deviation: 0.02E.V. Range: 2.20 to 2.68 % in Tolerance 100.00 %

% Bias 1.94 %% RSD 0.83 %

2.12.22.32.42.52.6

23

-Fe

b-2

01

1

19

-Ap

r-201

1

19

-Ap

r-201

1

19

-Ap

r-201

1

21

-Ap

r-201

1

07

-Ma

y-20

11

16

-Ma

y-20

11

16

-Ma

y-20

11

17

-Ma

y-20

11

17

-Ma

y-20

11

26

-Ma

y-20

11

08

-Jun-2

011

07

-Jul-201

1

08

-Jul-201

1

18

-Jul-201

1

18

-Jul-201

1

18

-Jul-201

1

20

-Jul-201

1

25

-Au

g-2

01

1

P_

Re

sult

(%

)

DateReported



-0.06

-0.04

-0.02

0.00

0.02

23

-Fe

b-2

01

1

19

-Ap

r-201

1

19

-Ap

r-201

1

19

-Ap

r-201

1

21

-Ap

r-201

1

07

-Ma

y-2

01

1

16

-Ma

y-2

01

1

16

-Ma

y-2

01

1

17

-Ma

y-2

01

1

17

-Ma

y-2

01

1

26

-Ma

y-2

01

1

08

-Jun

-20

11

07

-Jul-2

01

1

08

-Jul-2

01

1

18

-Jul-2

01

1

18

-Jul-2

01

1

18

-Jul-2

01

1

20

-Jul-2

01

1

25

-Au

g-2

01

1

Cum

ulat

ive S

um

of

P_R

esu

lt -

Me

an (

%)

DateReported

Cumulative Deviation from Assay Mean(Standard ST1000045 P ME-XRF12 Analysis )


0.0

0.2

0.4

0.6

0.8

1.0

23

-Fe

b-2

01

1

19

-Apr-2

01

1

19

-Apr-2

01

1

19

-Apr-2

01

1

21

-Apr-2

01

1

07

-Ma

y-2

01

1

16

-Ma

y-2

01

1

16

-Ma

y-2

01

1

17

-Ma

y-2

01

1

17

-Ma

y-2

01

1

26

-Ma

y-2

01

1

08

-Jun

-20

11

07

-Jul-2

01

1

08

-Jul-2

01

1

18

-Jul-2

01

1

18

-Jul-2

01

1

18

-Jul-2

01

1

20

-Jul-2

01

1

25

-Aug

-20

11

Cu

mula

tiv

e S

um

of

P_R

esu

lt -

Ex

pect

ed V

alu

e (

DateReported

Cumulative Deviation from Expected Value(Standard ST1000045 P ME-XRF12 Analysis )


Berlin Project Standards Feb 2012(Standard ST1000045 P ME-XRF12 Analysis )




Standard: AMIS-055 No of Analyses: 8Element: P_Result Minimum: 9.21Units: Maximum: 9.64Detection Limit: Mean: 9.36Expected Value (EV): 9.36 Std Deviation: 0.12E.V. Range: 8.42 to 10.30 % in Tolerance 100.00 %

% Bias 0.01 %% RSD 1.32 %

8.0

8.5

9.0

9.5

10 .0

10 .5

20

-Jul-2

01

1

19

-Aug

-20

11

26

-Aug

-20

11

05

-Se

p-2

01

1

19

-Se

p-2

01

1

01

-Oct-2

01

1

01

-Oct-2

01

1

P_

Resu

lt (

%)

DateReported

Standard Control Plot(Standard AMIS-055 P ME-MS81 Analysis)


-0.20

-0.10

0.00

0.10

20

-Jul-20

11

19

-Au

g-2

01

1

26

-Au

g-2

01

1

05

-Sep

-201

1

19

-Sep

-201

1

01

-Oct-2

01

1

01

-Oct-2

01

1

Cu

mula

tiv

e S

um

of

P_R

esu

lt -

Mean

(%

)

DateReported

Cumulative Deviation from Assay Mean(Standard AMIS-055 P ME-MS81 Analysis)


-0.20

-0.10

0.00

0.10

20

-Jul-2

01

1

19

-Aug

-20

11

26

-Aug

-20

11

05

-Se

p-2

01

1

19

-Se

p-2

01

1

01

-Oct-2

01

1

01

-Oct-2

01

1

Cu

mula

tiv

e S

um

of

P_R

esu

lt -

Ex

pe

cte

d V

alu

e (

DateReported

Cumulative Deviation from Expected Value(Standard AMIS-055 P ME-MS81 Analysis)

P_Result Mean of Cumulative Sum of P_Result - Expected Value (%) = -0.04

Berlin Project Standards Feb 2012(Standard AMIS-055 P ME-MS81 Analysis)




Standard: AMIS-055 No of Analyses: 9Element: P_Result Minimum: 9.21Units: Maximum: 9.64Detection Limit: Mean: 9.36Expected Value (EV): 9.36 Std Deviation: 0.12E.V. Range: 8.42 to 10.30 % in Tolerance 100.00 %

% Bias -0.02 %% RSD 1.25 %

8.0

8.5

9.0

9.5

10 .0

10 .5

18

-Aug

-20

11

19

-Aug

-20

11

26

-Aug

-20

11

05

-Se

p-2

01

1

19

-Se

p-2

01

1

19

-Se

p-2

01

1

01

-Oct-2

01

1

01

-Oct-2

01

1

P_

Resu

lt (

%)

DateReported

Standard Control Plot(Standard AMIS-055 P ME-XRF12 Analysis)


-0.20

-0.10

0.00

0.10

18

-Au

g-2

01

1

19

-Au

g-2

01

1

26

-Au

g-2

01

1

05

-Sep

-201

1

19

-Sep

-201

1

19

-Sep

-201

1

01

-Oct-2

01

1

01

-Oct-2

01

1

Cu

mula

tiv

e S

um

of

P_R

esu

lt -

Mean

(%

)

DateReported

Cumulative Deviation from Assay Mean(Standard AMIS-055 P ME-XRF12 Analysis)


-0.20

-0.10

0.00

0.10

18

-Aug

-20

11

19

-Aug

-20

11

26

-Aug

-20

11

05

-Se

p-2

01

1

19

-Se

p-2

01

1

19

-Se

p-2

01

1

01

-Oct-2

01

1

01

-Oct-2

01

1

Cu

mula

tiv

e S

um

of

P_R

esu

lt -

Ex

pe

cte

d V

alu

e (

DateReported

Cumulative Deviation from Expected Value(Standard AMIS-055 P ME-XRF12 Analysis)

P_Result Mean of Cumulative Sum of P_Result - Expected Value (%) = -0.05

Berlin Project Standards Feb 2012(Standard AMIS-055 P ME-XRF12 Analysis)


Berlin Project Yttrium Standards Feb 2012


Standard:AMIS-055 Y ME-MS81

Analysis No of Analyses: 14Element: Y Result Minimum: 51.20Units: ppm Maximum: 61.40Detection Limit: Mean: 56.29Expected Value (EV): 53.60 Std Deviation: 3.01E.V. Range: 48.24 to 58.96 % in Tolerance 71.43 %

% Bias 5.01 %% RSD 5.35 %

4850525456586062

20-Ju

l-20

11

30-Ju

l-20

11

04-A

ug-2

01

1

13-A

ug-2

01

1

18-A

ug-2

01

1

19-A

ug-2

01

1

26-A

ug-2

01

1

05-S

ep

-20

11

19-S

ep

-20

11

19-S

ep

-20

11

19-S

ep

-20

11

01-O

ct-20

11

01-O

ct-20

11

Y_

Res

ult

(ppm

)

DateReported

Standard Control Plot(Standard AMIS-055 ME-MS61U Analysis)

Y_Result Expected Value = 53.60 EV Range (48.24 to 58.96) Mean of Y_Result = 56.29

-15-10

-5

05

10

20-Ju

l-20

11

30-Ju

l-20

11

04-A

ug-2

01

1

13-A

ug-2

01

1

18-A

ug-2

01

1

19-A

ug-2

01

1

26-A

ug-2

01

1

05-S

ep

-20

11

19-S

ep

-20

11

19-S

ep

-20

11

19-S

ep

-20

11

01-O

ct-20

11

01-O

ct-20

11C

um

ula

tive

Su

m o

f Y

_R

esu

lt -

Mean

(p

pm

)

DateReported

Cumulative Deviation from Assay Mean(Standard AMIS-055 ME-MS61U Analysis)

Y_Result Mean of Cumulative Sum of Y_Result - Mean (ppm) = -0.97

-100

10

203040

20-Ju

l-20

11

30-Ju

l-20

11

04-A

ug-2

01

1

13-A

ug-2

01

1

18-A

ug-2

01

1

19-A

ug-2

01

1

26-A

ug-2

01

1

05-S

ep-2

01

1

19-S

ep-2

01

1

19-S

ep-2

01

1

19-S

ep-2

01

1

01-O

ct-20

11

01-O

ct-20

11

Cum

ula

tive

Su

m o

f Y

_R

esu

lt -

Exp

ect

ed V

alu

e (

DateReported

Cumulative Deviation from Expected Value(Standard AMIS-055 ME-MS61U Analysis)

Y_Result Mean of Cumulative Sum of Y_Result - Expected Value (ppm) = 19.17

Berlin Project Standards Feb 2012(Standard AMIS-055 ME-MS61U Analysis)

Printed: 02-Mar-2012 15:25:51 Data Imported: 03-Feb-2012 09:42:32 Page 1

Berlin Project Yttrium Standards Feb 2012


Standard: AMIS-055 No of Analyses: 8Element: Y Result Minimum: 51.50Units: ppm Maximum: 58.90Detection Limit: - Mean: 55.41Expected Value (EV): 53.60 Std Deviation: 2.52E.V. Range: 48.24 to 58.96 % in Tolerance 100.00 %

% Bias 3.38 %% RSD 4.54 %

48505254565860

20

-Jul-2

011

19

-Aug

-20

11

26

-Aug

-20

11

05

-Se

p-2

01

1

19

-Se

p-2

01

1

01

-Oct-2

01

1

01

-Oct-2

01

1

Y_

Re

sult

(p

pm

)

DateReported

Standard Control Plot(Standard AMIS-055 ME-MS81 Analysis )

Y_Result Expected Value = 53.60 EV Range (48.24 to 58.96) Mean of Y_Result = 55.41

-10123456

20

-Jul-201

1

19

-Au

g-2

01

1

26

-Au

g-2

01

1

05

-Sep

-2011

19

-Sep

-2011

01

-Oct-2

01

1

01

-Oct-2

01

1Cum

ula

tiv

e Su

m o

f Y_R

esu

lt -

Me

an (

pp

m)

DateReported

Cumulative Deviation from Assay Mean(Standard AMIS-055 ME-MS81 Analysis )

Y_Result Mean of Cumulative Sum of Y_Result - Mean (ppm) = 2.51

468

10121416

20

-Jul-2

01

1

19

-Aug

-20

11

26

-Aug

-20

11

05

-Se

p-2

01

1

19

-Se

p-2

01

1

01

-Oct-2

011

01

-Oct-2

011

Cu

mula

tiv

e S

um

of

Y_R

esu

lt -

Ex

pe

cted V

alu

e (

p

DateReported

Cumulative Deviation from Expected Value(Standard AMIS-055 ME-MS81 Analysis )

Y_Result Mean of Cumulative Sum of Y_Result - Expected Value (ppm) = 10.66

Berlin Project Standards Feb 2012(Standard AMIS-055 ME-MS81 Analysis )


Berlin Project U3O8 Corp. Blanks


Standard: Blank No of Analyses: 63Element: U Result Minimum: 0.90Units: ppm Maximum: 18.90Detection Limit: 1 Mean: 2.85Expected Value (EV): 2.97 Std Deviation: 2.89E.V. Range: 2.67 to 3.27 % in Tolerance 12.70 %

% Bias -3.96 %% RSD 101.36 %

0

5

10

15

20

10

-Ap

r-201

1

25

-Ma

y-20

11

02

-Jul-201

1

04

-Au

g-2

01

1

20

-Au

g-2

01

1

01

-Oct-2

01

1

U_R

esu

lt (

pp

m)

DateReported

Standard Control Plot(Blank ME-MS61U Analysis)

U_Result Expected Value = 2.97 EV Range (2.67 to 3.27) Mean of U_Result = 2.85

-60-50-40-30-20-10

0

10

-Ap

r-201

1

25

-Ma

y-2

01

1

02

-Jul-2

01

1

04

-Au

g-2

01

1

20

-Au

g-2

01

1

01

-Oct-2

01

1Cu

mula

tiv

e S

um

of

U_

Resu

lt -

Mea

n (

pp

m)

DateReported

Cumulative Deviation from Assay Mean(Blank ME-MS61U Analysis)


-60-50-40-30-20-10

0

10

-Apr-2

01

1

25

-Ma

y-2

01

1

02

-Jul-2

01

1

04

-Aug

-20

11

20

-Aug

-20

11

01

-Oct-2

01

1

Cu

mula

tiv

e S

um

of

U_

Resu

lt -

Ex

pect

ed

Va

lue (

p

DateReported

Cumulative Deviation from Expected Value(Blank ME-MS61U Analysis)


Berlin Project Standards Feb 2012(Blank ME-MS61U Analysis)


Berlin Project U3O8 Corp. Blanks


Standard: Blank No of Analyses: 24Element: U Result Minimum: 1.31Units: ppm Maximum: 11.65Detection Limit: 1 Mean: 3.52Expected Value (EV): 2.97 Std Deviation: 2.64E.V. Range: 2.67 to 3.27 % in Tolerance 0.00 %

% Bias 18.43 %% RSD 75.03 %

02468

1012

02

-De

c-201

0

10

-Jan

-20

11

11

-Jan

-20

11

02

-Fe

b-2

011

02

-Fe

b-2

011

04

-Fe

b-2

011

28

-Fe

b-2

011

04

-Ma

r-201

1

12

-Ma

r-201

1

09

-Ap

r-201

1

10

-Ap

r-201

1

13

-Ap

r-201

1

19

-Ap

r-201

1

21

-Ap

r-201

1

18

-Au

g-20

11

19

-Au

g-20

11

26

-Au

g-20

11

05

-Se

p-2

011

19

-Se

p-2

011

19

-Se

p-2

011

19

-Se

p-2

011

01-O

ct-20

11

01-O

ct-20

11

U_R

esu

lt (

pp

m)

DateReported

Standard Control Plot(Blank ME-MS81 Analysis )

U_Result Expected Value = 2.97 EV Range (2.67 to 3.27) Mean of U_Result = 3.52

-30

-20

-10

0

02

-Dec-2

01

0

10

-Jan

-20

11

11

-Jan

-20

11

02

-Fe

b-2

01

1

02

-Fe

b-2

01

1

04

-Fe

b-2

01

1

28

-Fe

b-2

01

1

04

-Ma

r-20

11

12

-Ma

r-20

11

09

-Apr-2

01

1

10

-Apr-2

01

1

13

-Apr-2

01

1

19

-Apr-2

01

1

21

-Apr-2

01

1

18

-Au

g-2

01

1

19

-Au

g-2

01

1

26

-Au

g-2

01

1

05

-Se

p-2

01

1

19

-Se

p-2

01

1

19

-Se

p-2

01

1

19

-Se

p-2

01

1

01

-Oct-2

01

1

01

-Oct-2

01

1Cu

mula

tiv

e S

um

of

U_

Resu

lt -

Mea

n (

pp

m)

DateReported

Cumulative Deviation from Assay Mean(Blank ME-MS81 Analysis )


-20

-10

0

10

20

02

-De

c-201

0

10

-Jan

-20

11

11

-Jan

-20

11

02

-Fe

b-2

01

1

02

-Fe

b-2

01

1

04

-Fe

b-2

01

1

28

-Fe

b-2

01

1

04

-Ma

r-20

11

12

-Ma

r-20

11

09

-Ap

r-201

1

10

-Ap

r-201

1

13

-Ap

r-201

1

19

-Ap

r-201

1

21

-Ap

r-201

1

18

-Au

g-2

01

1

19

-Au

g-2

01

1

26

-Au

g-2

01

1

05

-Se

p-2

01

1

19

-Se

p-2

01

1

19

-Se

p-2

01

1

19

-Se

p-2

01

1

01

-Oct-2

01

1

01

-Oct-2

01

1

Cu

mula

tiv

e S

um

of

U_

Resu

lt -

Ex

pect

ed V

alu

e (

p

DateReported

Cumulative Deviation from Expected Value(Blank ME-MS81 Analysis )


Berlin Project Standards Feb 2012(Blank ME-MS81 Analysis )


Berlin Project U3O8 Corp. Duplicates


Orig_Assay_U

Duplicate_Assay_U Units Result

No. Pairs: 4 4 Pearson CC: 1.00Minimum: 7.70 8.00 ppm Spearman CC: 1.00Maximum: 710.00 710.00 ppm Mean HARD: 0.77Mean: 243.68 243.00 ppm Median HARD: 0.59Median 128.50 127.00 ppmStd. Deviation: 273.88 274.13 ppm Mean HRD: -0.18Coefficient of Variation: 1.12 1.13 Median HRD 0.27

0

5

10

1 10 10 0 10 00

HA

RD

(%

)

Mean of Data Pair (ppm)

Mean vs. HARD Plot(Trench Lab Pulp U ME-MS61 Analysis)

Mean HARD: 0.77 Median HARD: 0.59Precision: 10%

0

50

10 0

0 10 20 30 40 50 60 70 80 90 10 0H

AR

D (

%)

Rank (%)

Rank HARD Plot(Trench Lab Pulp U ME-MS61 Analysis)

Precision: 10%

0

50

10 0

-1.0 0.0 1.0Fre

qu

en

cy (

%)

HRD (/100)

HRD Histogram(Trench Lab Pulp U ME-MS61 Analysis)

Mean HRD: -0.18 Median HRD: 0.27Precision: +/-10%

-10-505

10

1 10 10 0 10 00

HR

D (

%)


Mean vs. HRD Plot(Trench Lab Pulp U ME-MS61 Analysis)


0.1

1

10

10 0

10 00

1 10 10 0 10 00

Ab

solu

te D

iffe

ren

ce (

ppm

)


T & H Precision Plot (Assay Pairs)(Trench Lab Pulp U ME-MS61 Analysis)

10% 20% 30% 50%

T & H Precision Plot (Grouped Pairs)(Trench Lab Pulp U ME-MS61 Analysis)

There is not enough data to generate this plot.

0

20 0

40 0

60 0

80 0

0 10 0 20 0 30 0 40 0 50 0 60 0 70 0 80 0

Du

plica

te_

Ass

ay_

U (

pp

m)

Orig_Assay_U (ppm)

Correlation Plot(Trench Lab Pulp U ME-MS61 Analysis)

P.CC= 1.00 S.CC= 1.00 Ref. Line y = 1.00x -0.90

0

20 0

40 0

60 0

80 0

0 10 0 20 0 30 0 40 0 50 0 60 0 70 0 80 0

Du

plica

te_

Ass

ay_

U (

pp

m)

Orig_Assay_U (ppm)

QQ Plot(Trench Lab Pulp U ME-MS61 Analysis)

Ref. Line y = 1.00x -0.90

Berlin Project Duplicate Analysis Feb 12(Trench Lab Pulp U ME-MS61 Analysis)




Orig_Assay_U



0

5

10

1 10 10 0 10 00

HA

RD

(%

)


Mean vs. HARD Plot(Lab Pulp U ME-MS61 Analysis)


0

50

10 0

0 10 20 30 40 50 60 70 80 90 10 0H

AR

D (

%)

Rank (%)

Rank HARD Plot(Lab Pulp U ME-MS61 Analysis)

Precision: 10%

0

50

10 0

-1.0 0.0 1.0Fre

qu

en

cy (

%)

HRD (/100)

HRD Histogram(Lab Pulp U ME-MS61 Analysis)


-10-505

10

1 10 10 0 10 00

HR

D (

%)


Mean vs. HRD Plot(Lab Pulp U ME-MS61 Analysis)


0.1

1

10

10 0

10 00

1 10 10 0 10 00

Ab

solu

te D

iffe

ren

ce (

pp

m)


T & H Precision Plot (Assay Pairs)(Lab Pulp U ME-MS61 Analysis)

10% 20% 30% 50%

T & H Precision Plot (Grouped Pairs)(Lab Pulp U ME-MS61 Analysis)


0

20 0

40 0

60 0

80 0

0 10 0 20 0 30 0 40 0 50 0 60 0 70 0 80 0

Du

plica

te_

Ass

ay_

U (

ppm

)

Orig_Assay_U (ppm)

Correlation Plot(Lab Pulp U ME-MS61 Analysis)


0

20 0

40 0

60 0

80 0

0 10 0 20 0 30 0 40 0 50 0 60 0 70 0 80 0

Du

plica

te_

Ass

ay_

U (

ppm

)

Orig_Assay_U (ppm)

QQ Plot(Lab Pulp U ME-MS61 Analysis)

Ref. Line y = 1.00x -0.90

Berlin Project Duplicate Analysis Feb 12(Lab Pulp U ME-MS61 Analysis)




Orig_Assay_U



0

20

40

60

1 10 10 0 10 00

HA

RD

(%

)


Mean vs. HARD Plot(U3O8Corp 10 Mesh U ME-MS61U Analysis)


0

50

10 0

0 10 20 30 40 50 60 70 80 90 10 0H

AR

D (

%)

Rank (%)

Rank HARD Plot(U3O8Corp 10 Mesh U ME-MS61U Analysis)

Precision: 10%

80.00% of data are withinPrecision limits

020406080

-1.0 0.0 1.0Fre

qu

en

cy (

%)

HRD (/100)

HRD Histogram(U3O8Corp 10 Mesh U ME-MS61U Analysis)


-60-40-20

020

1 10 10 0 10 00

HR

D (

%)


Mean vs. HRD Plot(U3O8Corp 10 Mesh U ME-MS61U Analysis)


0.1

1

10

10 0

10 00

1 10 10 0 10 00

Ab

solu

te D

iffe

ren

ce (

pp

m)


T & H Precision Plot (Assay Pairs)(U3O8Corp 10 Mesh U ME-MS61U Analysis)

10% 20% 30% 50%

T & H Precision Plot (Grouped Pairs)(U3O8Corp 10 Mesh U ME-MS61U Analysis)


020 040 060 080 0

0 10 0 20 0 30 0 40 0 50 0 60 0 70 0 80 0

Du

plica

te_

Ass

ay_

U (

ppm

)

Orig_Assay_U (ppm)

Correlation Plot(U3O8Corp 10 Mesh U ME-MS61U Analysis)

P.CC= 1.00 S.CC= 0.93 Ref. Liney = 0.98x + 2.51

0

20 0

40 0

60 0

80 0

0 10 0 20 0 30 0 40 0 50 0 60 0 70 0 80 0

Du

plica

te_

Ass

ay_

U (

ppm

)

Orig_Assay_U (ppm)

QQ Plot(U3O8Corp 10 Mesh U ME-MS61U Analysis)

Ref. Line y = 0.98x + 2.50

Berlin Project Duplicate Analysis Feb 12(U3O8Corp 10 Mesh U ME-MS61U Analysis)




Orig_Assay_U


No. Pairs: 15 15 Pearson CC: 1.00Minimum: 10.00 5.90 ppm Spearman CC: 0.99Maximum: 1,010.00 970.00 ppm Mean HARD: 4.35Mean: 247.67 248.28 ppm Median HARD: 2.40Median 27.10 23.30 ppmStd. Deviation: 339.81 343.76 ppm Mean HRD: 1.69Coefficient of Variation: 1.37 1.38 Median HRD 0.81

0

10

20

30

1 10 10 0 10 00

HA

RD

(%

)


Mean vs. HARD Plot(U3O8Corp Sample DH U ME-MS61U Analysis)


0

50

10 0

0 10 20 30 40 50 60 70 80 90 10 0

HA

RD

(%

)Rank (%)

Rank HARD Plot(U3O8Corp Sample DH U ME-MS61U Analysis)

Precision: 10%


0

20

40

60

-1.0 0.0 1.0Fre

qu

en

cy (

%)

HRD (/100)

HRD Histogram(U3O8Corp Sample DH U ME-MS61U Analysis)

Mean HRD: 1.69 Median HRD: 0.81Precision: +/-10%

-100

102030

1 10 10 0 10 00

HR

D (

%)


Mean vs. HRD Plot(U3O8Corp Sample DH U ME-MS61U Analysis)


0.1

1

10

10 0

10 00

1 10 10 0 10 00

Ab

solu

te D

iffe

ren

ce (

pp

m)


T & H Precision Plot (Assay Pairs)(U3O8Corp Sample DH U ME-MS61U Analysis)

10% 20% 30% 50%

1

10

10 0

10 10 0Med

ian

AD

(p

pm

)

Grouped Mean of Pair (ppm)

T & H Precision Plot (Grouped Pairs)(U3O8Corp Sample DH U ME-MS61U Analysis)

10% 20% 30% 50%

0

50 0

10 00

15 00

0 20 0 40 0 60 0 80 0 10 00 12 00 14 00

Du

plica

te_

Ass

ay_

U (

ppm

)

Orig_Assay_U (ppm)

Correlation Plot(U3O8Corp Sample DH U ME-MS61U Analysis)


0

50 0

10 00

15 00

0 20 0 40 0 60 0 80 0 10 00 12 00 14 00

Du

plica

te_

Ass

ay_

U (

ppm

)

Orig_Assay_U (ppm)

QQ Plot(U3O8Corp Sample DH U ME-MS61U Analysis)

Ref. Line y = 1.01x -1.65

Berlin Project Duplicate Analysis Feb 12(U3O8Corp Sample DH U ME-MS61U Analysis)




Orig_Assay_U


No. Pairs: 24 24 Pearson CC: 1.00Minimum: 5.60 4.30 ppm Spearman CC: 0.99Maximum: 1,730.00 1,760.00 ppm Mean HARD: 2.46Mean: 231.33 230.47 ppm Median HARD: 2.02Median 15.90 16.30 ppmStd. Deviation: 447.39 448.97 ppm Mean HRD: 0.16Coefficient of Variation: 1.93 1.95 Median HRD -0.65

0

5

10

15

1 10 10 0 10 00 10 000

HA

RD

(%

)


Mean vs. HARD Plot(Lab Pulp U ME-MS61U Analysis )


0

50

10 0

0 10 20 30 40 50 60 70 80 90 10 0H

AR

D (

%)

Rank (%)

Rank HARD Plot(Lab Pulp U ME-MS61U Analysis )

Precision: 10%


020406080

-1.0 0.0 1.0Fre

qu

en

cy (

%)

HRD (/100)

HRD Histogram(Lab Pulp U ME-MS61U Analysis )

Mean HRD: 0.16 Median HRD: -0.65Precision: +/-10%

-10

0

10

20

1 10 10 0 10 00 10 000

HR

D (

%)


Mean vs. HRD Plot(Lab Pulp U ME-MS61U Analysis )


0.1

1

10

10 0

10 00

1 10 10 0 10 00 10 000

Ab

solu

te D

iffe

ren

ce (

ppm

)


T & H Precision Plot (Assay Pairs)(Lab Pulp U ME-MS61U Analysis )

10% 20% 30% 50%

0.1

1

10

10 0

10 00

1 10 10 0 10 00Media

n A

D (

ppm

)


T & H Precision Plot (Grouped Pairs)(Lab Pulp U ME-MS61U Analysis )

10% 20% 30% 50%

0

50 0

10 00

15 00

20 00

0 50 0 10 00 15 00 20 00

Du

plica

te_

Ass

ay_

U (

pp

m)

Orig_Assay_U (ppm)

Correlation Plot(Lab Pulp U ME-MS61U Analysis )


0

50 0

10 00

15 00

20 00

0 50 0 10 00 15 00 20 00

Du

plica

te_

Ass

ay_

U (

pp

m)

Orig_Assay_U (ppm)

QQ Plot(Lab Pulp U ME-MS61U Analysis )

Ref. Line y = 1.00x -1.56

Berlin Project Duplicate Analysis Feb 12(Lab Pulp U ME-MS61U Analysis )




Orig_Assay_U


No. Pairs: 5 5 Pearson CC: 1.00Minimum: 8.64 8.42 ppm Spearman CC: 1.00Maximum: 788.00 838.00 ppm Mean HARD: 1.91Mean: 295.89 302.32 ppm Median HARD: 1.98Median 258.00 248.00 ppmStd. Deviation: 272.99 292.17 ppm Mean HRD: 0.68Coefficient of Variation: 0.92 0.97 Median HRD 1.29

0

5

10

1 10 10 0 10 00

HA

RD

(%

)


Mean vs. HARD Plot(U3O8Corp 10 Mesh U ME-MS81 Analysis)


0

50

10 0

0 10 20 30 40 50 60 70 80 90 10 0H

AR

D (

%)

Rank (%)

Rank HARD Plot(U3O8Corp 10 Mesh U ME-MS81 Analysis)

Precision: 10%

0

20

40

60

-1.0 0.0 1.0Fre

qu

en

cy (

%)

HRD (/100)

HRD Histogram(U3O8Corp 10 Mesh U ME-MS81 Analysis)


-10-505

10

1 10 10 0 10 00

HR

D (

%)


Mean vs. HRD Plot(U3O8Corp 10 Mesh U ME-MS81 Analysis)


0.1

1

10

10 0

10 00

1 10 10 0 10 00

Abso

lute

Dif

fere

nce

(pp

m)


T & H Precision Plot (Assay Pairs)(U3O8Corp 10 Mesh U ME-MS81 Analysis)

10% 20% 30% 50%

T & H Precision Plot (Grouped Pairs)(U3O8Corp 10 Mesh U ME-MS81 Analysis)


-50 0

0

50 0

10 00

0 20 0 40 0 60 0 80 0 10 00

Du

plica

te_

Ass

ay_

U (

pp

m)

Orig_Assay_U (ppm)

Correlation Plot(U3O8Corp 10 Mesh U ME-MS81 Analysis)

P.CC= 1.00 S.CC= 1.00 Ref. Liney = 1.07x -14.12

-50 0

0

50 0

10 00

0 20 0 40 0 60 0 80 0 10 00

Du

plica

te_

Ass

ay_

U (

ppm

)

Orig_Assay_U (ppm)

QQ Plot(U3O8Corp 10 Mesh U ME-MS81 Analysis)

Ref. Line y = 1.07x -14.12

Berlin Project Duplicate Analysis Feb 12(U3O8Corp 10 Mesh U ME-MS81 Analysis)




Orig_Assay_U



0

5

10

1 10 10 0 10 00

HA

RD

(%

)


Mean vs. HARD Plot(Trench Pulp Duplicate U ME-MS81 Analysis )


0

50

10 0

0 10 20 30 40 50 60 70 80 90 10 0H

AR

D (

%)

Rank (%)

Rank HARD Plot(Trench Pulp Duplicate U ME-MS81 Analysis )

Precision: 10%

0

50

10 0

-1.0 0.0 1.0Fre

qu

en

cy (

%)

HRD (/100)

HRD Histogram(Trench Pulp Duplicate U ME-MS81 Analysis )


-10-505

10

1 10 10 0 10 00

HR

D (

%)


Mean vs. HRD Plot(Trench Pulp Duplicate U ME-MS81 Analysis )


0.1

1

10

10 0

10 00

10 10 0 10 00

Ab

solu

te D

iffe

ren

ce (

pp

m)


T & H Precision Plot (Assay Pairs)(Trench Pulp Duplicate U ME-MS81 Analysis )

10% 20% 30% 50%

1

10

10 0

10 0 10 00Med

ian

AD

(p

pm

)


T & H Precision Plot (Grouped Pairs)(Trench Pulp Duplicate U ME-MS81 Analysis )

10% 20% 30% 50%

020 040 060 080 0

0 10 0 20 0 30 0 40 0 50 0 60 0 70 0 80 0

Du

plica

te_

Ass

ay_

U (

ppm

)

Orig_Assay_U (ppm)

Correlation Plot(Trench Pulp Duplicate U ME-MS81 Analysis )


0

20 0

40 0

60 0

80 0

0 10 0 20 0 30 0 40 0 50 0 60 0 70 0 80 0

Du

plica

te_

Ass

ay_

U (

ppm

)

Orig_Assay_U (ppm)

QQ Plot(Trench Pulp Duplicate U ME-MS81 Analysis )

Ref. Line y = 0.99x + 0.24

Berlin Project Duplicate Statistics Feb 12(Trench Pulp Duplicate U ME-MS81 Analysis )




Orig_Assay_U


No. Pairs: 8 8 Pearson CC: 1.00Minimum: 21.10 22.20 ppm Spearman CC: 1.00Maximum: 730.00 793.00 ppm Mean HARD: 2.04Mean: 318.23 331.94 ppm Median HARD: 1.47Median 285.75 287.50 ppmStd. Deviation: 256.80 271.59 ppm Mean HRD: -1.55Coefficient of Variation: 0.81 0.82 Median HRD -0.67

0

5

10

1 10 10 0 10 00

HA

RD

(%

)


Mean vs. HARD Plot(U3O8Corp DH Duplicate U ME-MS81 Analysis)


0

50

10 0

0 10 20 30 40 50 60 70 80 90 10 0

HA

RD

(%

)Rank (%)

Rank HARD Plot(U3O8Corp DH Duplicate U ME-MS81 Analysis)

Precision: 10%

020406080

-1.0 0.0 1.0Fre

qu

en

cy (

%)

HRD (/100)

HRD Histogram(U3O8Corp DH Duplicate U ME-MS81 Analysis)

Mean HRD: -1.55 Median HRD: -0.67Precision: +/-10%

-10-505

10

1 10 10 0 10 00

HR

D (

%)


Mean vs. HRD Plot(U3O8Corp DH Duplicate U ME-MS81 Analysis)


0.1

1

10

10 0

10 00

10 10 0 10 00

Ab

solu

te D

iffe

ren

ce (

ppm

)


T & H Precision Plot (Assay Pairs)(U3O8Corp DH Duplicate U ME-MS81 Analysis)

10% 20% 30% 50%

T & H Precision Plot (Grouped Pairs)(U3O8Corp DH Duplicate U ME-MS81 Analysis)


0

20 0

40 0

60 0

80 0

0 10 0 20 0 30 0 40 0 50 0 60 0 70 0 80 0

Du

plica

te_

Ass

ay_

U (

pp

m)

Orig_Assay_U (ppm)

Correlation Plot(U3O8Corp DH Duplicate U ME-MS81 Analysis)


0

20 0

40 0

60 0

80 0

0 10 0 20 0 30 0 40 0 50 0 60 0 70 0 80 0

Du

plica

te_

Ass

ay_

U (

pp

m)

Orig_Assay_U (ppm)

QQ Plot(U3O8Corp DH Duplicate U ME-MS81 Analysis)

Ref. Line y = 1.05x -3.07

Berlin Project Duplicate Analysis Feb 12(U3O8Corp DH Duplicate U ME-MS81 Analysis)




Assay_V_Duplicate_A

ssay_V Units ResultNo. Pairs: 8 8 Pearson CC: 0.99Minimum: 152.00 149.00 ppm Spearman CC: 0.98Maximum: 3,320.00 3,410.00 ppm Mean HARD: 2.25Mean: 1,898.63 1,939.25 ppm Median HARD: 0.77Median 2,145.00 2,025.00 ppmStd. Deviation: 1,059.61 1,107.37 ppm Mean HRD: -0.47Coefficient of Variation: 0.56 0.57 Median HRD 0.29

0

5

10

1 10 10 0 10 00 10 000

HA

RD

(%

)


Mean vs. HARD Plot(DH Pulp Duplicate U ME-MS81 Analysis)


0

50

10 0

0 10 20 30 40 50 60 70 80 90 10 0H

AR

D (

%)

Rank (%)

Rank HARD Plot(DH Pulp Duplicate U ME-MS81 Analysis)

Precision: 10%

020406080

-1.0 0.0 1.0Fre

qu

en

cy (

%)

HRD (/100)

HRD Histogram(DH Pulp Duplicate U ME-MS81 Analysis)


-10-505

10

1 10 10 0 10 00 10 000

HR

D (

%)


Mean vs. HRD Plot(DH Pulp Duplicate U ME-MS81 Analysis)


1

10

10 0

10 00

10 000

10 0 10 00 10 000

Ab

solu

te D

iffe

ren

ce (

ppm

)


T & H Precision Plot (Assay Pairs)(DH Pulp Duplicate U ME-MS81 Analysis)

10% 20% 30% 50%

T & H Precision Plot (Grouped Pairs)(DH Pulp Duplicate U ME-MS81 Analysis)


010 0020 0030 0040 00

0 10 00 20 00 30 00 40 00

Du

plica

te_

Ass

ay_

V (

pp

m)

Assay_V_ (ppm)

Correlation Plot(DH Pulp Duplicate U ME-MS81 Analysis)


0

10 00

20 00

30 00

40 00

0 10 00 20 00 30 00 40 00

Du

plica

te_

Ass

ay_

V (

pp

m)

Assay_V_ (ppm)

QQ Plot(DH Pulp Duplicate U ME-MS81 Analysis)

Ref. Line y = 1.04x -28.23

Berlin Project Duplicate Statistics Feb 12(DH Pulp Duplicate U ME-MS81 Analysis)




Orig_Assay_U



0

5

10

1 10 10 0 10 00

HA

RD

(%

)


Mean vs. HARD Plot(Lab Pulp Duplicate U ME-MS81 Analysis )


0

50

10 0

0 10 20 30 40 50 60 70 80 90 10 0H

AR

D (

%)

Rank (%)

Rank HARD Plot(Lab Pulp Duplicate U ME-MS81 Analysis )

Precision: 10%

0

50

10 0

-1.0 0.0 1.0Fre

qu

en

cy (

%)

HRD (/100)

HRD Histogram(Lab Pulp Duplicate U ME-MS81 Analysis )


-10-505

10

1 10 10 0 10 00

HR

D (

%)


Mean vs. HRD Plot(Lab Pulp Duplicate U ME-MS81 Analysis )


0.1

1

10

10 0

10 00

10 10 0 10 00

Ab

solu

te D

iffe

ren

ce (

ppm

)


T & H Precision Plot (Assay Pairs)(Lab Pulp Duplicate U ME-MS81 Analysis )

10% 20% 30% 50%

T & H Precision Plot (Grouped Pairs)(Lab Pulp Duplicate U ME-MS81 Analysis )


0

50 0

10 00

0 20 0 40 0 60 0 80 0 10 00

Du

plica

te_

Ass

ay_

U (

pp

m)

Orig_Assay_U (ppm)

Correlation Plot(Lab Pulp Duplicate U ME-MS81 Analysis )


0

50 0

10 00

0 20 0 40 0 60 0 80 0 10 00

Du

plica

te_

Ass

ay_

U (

pp

m)

Orig_Assay_U (ppm)

QQ Plot(Lab Pulp Duplicate U ME-MS81 Analysis )

Ref. Line y = 1.00x -1.21

Berlin Project Duplicate Statistics Feb 12(Lab Pulp Duplicate U ME-MS81 Analysis )




Assay_P_Duplicate_A

ssay_P Units ResultNo. Pairs: 10 10 Pearson CC: 1.00Minimum: 0.00 0.00 ppm Spearman CC: 0.96Maximum: 3,740.00 3,670.00 ppm Mean HARD: 2.15Mean: 782.80 764.20 ppm Median HARD: 0.64Median 155.50 135.00 ppmStd. Deviation: 1,297.20 1,280.84 ppm Mean HRD: 1.87Coefficient of Variation: 1.66 1.68 Median HRD 0.17

0

5

10

1 10 10 0 10 00 10 000

HA

RD

(%

)


Mean vs. HARD Plot(Lab Pulp Duplicate P ME-MS81 Analysis )


0

50

10 0

0 10 20 30 40 50 60 70 80 90 10 0H

AR

D (

%)

Rank (%)

Rank HARD Plot(Lab Pulp Duplicate P ME-MS81 Analysis )

Precision: 10%

020406080

-1.0 0.0 1.0Fre

qu

en

cy (

%)

HRD (/100)

HRD Histogram(Lab Pulp Duplicate P ME-MS81 Analysis )


-10-505

10

1 10 10 0 10 00 10 000

HR

D (

%)


Mean vs. HRD Plot(Lab Pulp Duplicate P ME-MS81 Analysis )


1

10

10 0

10 00

10 000

10 10 0 10 00 10 000

Ab

solu

te D

iffe

ren

ce (

pp

m)


T & H Precision Plot (Assay Pairs)(Lab Pulp Duplicate P ME-MS81 Analysis )

10% 20% 30% 50%

T & H Precision Plot (Grouped Pairs)(Lab Pulp Duplicate P ME-MS81 Analysis )


-20 00

0

20 00

40 00

0 10 00 20 00 30 00 40 00

Du

plica

te_

Ass

ay_

P (

ppm

)

Assay_P_ (ppm)

Correlation Plot(Lab Pulp Duplicate P ME-MS81 Analysis )


-20 00

0

20 00

40 00

0 10 00 20 00 30 00 40 00

Du

plica

te_

Ass

ay_

P (

ppm

)

Assay_P_ (ppm)

QQ Plot(Lab Pulp Duplicate P ME-MS81 Analysis )

Ref. Line y = 0.99x -8.64

Berlin Project Duplicate Statistics Feb 12(Lab Pulp Duplicate P ME-MS81 Analysis )




Assay_V_Duplicate_A


0

5

10

1 10 10 0 10 00 10 000

HA

RD

(%

)


Mean vs. HARD Plot(Trench Pulp Duplicate V ME-MS81 Analysis )


0

50

10 0

0 10 20 30 40 50 60 70 80 90 10 0H

AR

D (

%)

Rank (%)

Rank HARD Plot(Trench Pulp Duplicate V ME-MS81 Analysis )

Precision: 10%

0

50

10 0

-1.0 0.0 1.0Fre

qu

en

cy (

%)

HRD (/100)

HRD Histogram(Trench Pulp Duplicate V ME-MS81 Analysis )


-10-505

10

1 10 10 0 10 00 10 000

HR

D (

%)


Mean vs. HRD Plot(Trench Pulp Duplicate V ME-MS81 Analysis )


1

10

10 0

10 00

10 000

10 0 10 00 10 000

Ab

solu

te D

iffe

ren

ce (

pp

m)


T & H Precision Plot (Assay Pairs)(Trench Pulp Duplicate V ME-MS81 Analysis )

10% 20% 30% 50%

10

10 0

10 00

10 00 10 000Med

ian

AD

(p

pm

)


T & H Precision Plot (Grouped Pairs)(Trench Pulp Duplicate V ME-MS81 Analysis )

10% 20% 30% 50%

0

50 00

10 000

0 20 00 40 00 60 00 80 00 10 000

Du

plica

te_

Ass

ay_

V (

ppm

)

Assay_V_ (ppm)

Correlation Plot(Trench Pulp Duplicate V ME-MS81 Analysis )


0

50 00

10 000

0 20 00 40 00 60 00 80 00 10 000

Du

plica

te_

Ass

ay_

V (

ppm

)

Assay_V_ (ppm)

QQ Plot(Trench Pulp Duplicate V ME-MS81 Analysis )

Ref. Line y = 1.00x -4.69

Berlin Project Duplicate Statistics Feb 12(Trench Pulp Duplicate V ME-MS81 Analysis )




Assay_V_Duplicate_A


0

5

10

1 10 10 0 10 00 10 000

HA

RD

(%

)


Mean vs. HARD Plot(DH Pulp Duplicate V ME-MS81 Analysis )


0

50

10 0

0 10 20 30 40 50 60 70 80 90 10 0H

AR

D (

%)

Rank (%)

Rank HARD Plot(DH Pulp Duplicate V ME-MS81 Analysis )

Precision: 10%

020406080

-1.0 0.0 1.0Fre

qu

en

cy (

%)

HRD (/100)

HRD Histogram(DH Pulp Duplicate V ME-MS81 Analysis )


-10-505

10

1 10 10 0 10 00 10 000

HR

D (

%)


Mean vs. HRD Plot(DH Pulp Duplicate V ME-MS81 Analysis )


1

10

10 0

10 00

10 000

10 0 10 00 10 000

Ab

solu

te D

iffe

ren

ce (

ppm

)


T & H Precision Plot (Assay Pairs)(DH Pulp Duplicate V ME-MS81 Analysis )

10% 20% 30% 50%

T & H Precision Plot (Grouped Pairs)(DH Pulp Duplicate V ME-MS81 Analysis )


010 0020 0030 0040 00

0 10 00 20 00 30 00 40 00

Du

plica

te_

Ass

ay_

V (

pp

m)

Assay_V_ (ppm)

Correlation Plot(DH Pulp Duplicate V ME-MS81 Analysis )


0

10 00

20 00

30 00

40 00

0 10 00 20 00 30 00 40 00

Du

plica

te_

Ass

ay_

V (

pp

m)

Assay_V_ (ppm)

QQ Plot(DH Pulp Duplicate V ME-MS81 Analysis )

Ref. Line y = 1.04x -28.23

Berlin Project Duplicate Statistics Feb 12(DH Pulp Duplicate V ME-MS81 Analysis )




Assay_V_Duplicate_A

ssay_V Units ResultNo. Pairs: 10 10 Pearson CC: 0.99Minimum: 1,260.00 1,280.00 ppm Spearman CC: 0.96Maximum: 5,980.00 5,610.00 ppm Mean HARD: 1.99Mean: 2,889.00 2,876.00 ppm Median HARD: 1.19Median 2,310.00 2,325.00 ppmStd. Deviation: 1,320.31 1,277.94 ppm Mean HRD: 0.21Coefficient of Variation: 0.46 0.44 Median HRD 0.57

0

5

10

1 10 10 0 10 00 10 000

HA

RD

(%

)


Mean vs. HARD Plot(Lab Pulp Duplicate V ME-MS81 Analysis )


0

50

10 0

0 10 20 30 40 50 60 70 80 90 10 0H

AR

D (

%)

Rank (%)

Rank HARD Plot(Lab Pulp Duplicate V ME-MS81 Analysis )

Precision: 10%

020406080

-1.0 0.0 1.0Fre

qu

en

cy (

%)

HRD (/100)

HRD Histogram(Lab Pulp Duplicate V ME-MS81 Analysis )


-10-505

10

1 10 10 0 10 00 10 000

HR

D (

%)


Mean vs. HRD Plot(Lab Pulp Duplicate V ME-MS81 Analysis )


10

10 0

10 00

10 000

10 00 10 000

Ab

solu

te D

iffe

ren

ce (

ppm

)


T & H Precision Plot (Assay Pairs)(Lab Pulp Duplicate V ME-MS81 Analysis )

10% 20% 30% 50%

T & H Precision Plot (Grouped Pairs)(Lab Pulp Duplicate V ME-MS81 Analysis )


0

20 00

40 00

60 00

0 10 00 20 00 30 00 40 00 50 00 60 00

Du

plica

te_

Ass

ay_

V (

pp

m)

Assay_V_ (ppm)

Correlation Plot(Lab Pulp Duplicate V ME-MS81 Analysis )


0

20 00

40 00

60 00

0 10 00 20 00 30 00 40 00 50 00 60 00

Du

plica

te_

Ass

ay_

V (

pp

m)

Assay_V_ (ppm)

QQ Plot(Lab Pulp Duplicate V ME-MS81 Analysis )

Ref. Line y = 0.96x + 99.17

Berlin Project Duplicate Statistics Feb 12(Lab Pulp Duplicate V ME-MS81 Analysis )




Assay_Y_Duplicate_A

ssay_Y Units ResultNo. Pairs: 12 12 Pearson CC: 1.00Minimum: 33.80 33.60 ppm Spearman CC: 1.00Maximum: 601.00 623.00 ppm Mean HARD: 0.94Mean: 221.64 222.18 ppm Median HARD: 0.72Median 126.00 126.00 ppmStd. Deviation: 193.84 198.00 ppm Mean HRD: 0.21Coefficient of Variation: 0.87 0.89 Median HRD 0.44

0

5

10

1 10 10 0 10 00

HA

RD

(%

)


Mean vs. HARD Plot(Trench Pulp Duplicate Y ME-MS81 Analysis )


0

50

10 0

0 10 20 30 40 50 60 70 80 90 10 0H

AR

D (

%)

Rank (%)

Rank HARD Plot(Trench Pulp Duplicate Y ME-MS81 Analysis )

Precision: 10%

0

50

10 0

-1.0 0.0 1.0Fre

qu

en

cy (

%)

HRD (/100)

HRD Histogram(Trench Pulp Duplicate Y ME-MS81 Analysis )


-10-505

10

1 10 10 0 10 00

HR

D (

%)


Mean vs. HRD Plot(Trench Pulp Duplicate Y ME-MS81 Analysis )


0.1

1

10

10 0

10 00

10 10 0 10 00

Ab

solu

te D

iffe

ren

ce (

pp

m)


T & H Precision Plot (Assay Pairs)(Trench Pulp Duplicate Y ME-MS81 Analysis )

10% 20% 30% 50%

1

10

10 0

10 0 10 00Med

ian

AD

(p

pm

)


T & H Precision Plot (Grouped Pairs)(Trench Pulp Duplicate Y ME-MS81 Analysis )

10% 20% 30% 50%

0

20 0

40 0

60 0

80 0

0 10 0 20 0 30 0 40 0 50 0 60 0 70 0 80 0

Du

plica

te_

Ass

ay_

Y (

pp

m)

Assay_Y_ (ppm)

Correlation Plot(Trench Pulp Duplicate Y ME-MS81 Analysis )


0

20 0

40 0

60 0

80 0

0 10 0 20 0 30 0 40 0 50 0 60 0 70 0 80 0

Du

plica

te_

Ass

ay_

Y (

pp

m)

Assay_Y_ (ppm)

QQ Plot(Trench Pulp Duplicate Y ME-MS81 Analysis )

Ref. Line y = 1.02x -4.11

Berlin Project Duplicate Statistics Feb 12(Trench Pulp Duplicate Y ME-MS81 Analysis )




Assay_Y_Duplicate_A

ssay_Y Units ResultNo. Pairs: 10 10 Pearson CC: 1.00Minimum: 14.30 14.10 ppm Spearman CC: 1.00Maximum: 831.00 837.00 ppm Mean HARD: 0.98Mean: 342.39 343.45 ppm Median HARD: 0.87Median 309.50 300.00 ppmStd. Deviation: 299.72 302.94 ppm Mean HRD: 0.33Coefficient of Variation: 0.88 0.88 Median HRD -0.11

0

5

10

1 10 10 0 10 00

HA

RD

(%

)


Mean vs. HARD Plot(Lab Pulp Duplicate Y ME-MS81 Analysis )


0

50

10 0

0 10 20 30 40 50 60 70 80 90 10 0H

AR

D (

%)

Rank (%)

Rank HARD Plot(Lab Pulp Duplicate Y ME-MS81 Analysis )

Precision: 10%

0

50

10 0

-1.0 0.0 1.0Fre

qu

en

cy (

%)

HRD (/100)

HRD Histogram(Lab Pulp Duplicate Y ME-MS81 Analysis )


-10-505

10

1 10 10 0 10 00

HR

D (

%)


Mean vs. HRD Plot(Lab Pulp Duplicate Y ME-MS81 Analysis )


0.1

1

10

10 0

10 00

10 10 0 10 00

Ab

solu

te D

iffe

ren

ce (

pp

m)


T & H Precision Plot (Assay Pairs)(Lab Pulp Duplicate Y ME-MS81 Analysis )

10% 20% 30% 50%

T & H Precision Plot (Grouped Pairs)(Lab Pulp Duplicate Y ME-MS81 Analysis )


0

50 0

10 00

0 20 0 40 0 60 0 80 0 10 00

Du

plica

te_

Ass

ay_

Y (

ppm

)

Assay_Y_ (ppm)

Correlation Plot(Lab Pulp Duplicate Y ME-MS81 Analysis )


0

50 0

10 00

0 20 0 40 0 60 0 80 0 10 00

Du

plica

te_

Ass

ay_

Y (

ppm

)

Assay_Y_ (ppm)

QQ Plot(Lab Pulp Duplicate Y ME-MS81 Analysis )

Ref. Line y = 1.01x -2.57

Berlin Project Duplicate Statistics Feb 12(Lab Pulp Duplicate Y ME-MS81 Analysis )




Assay_Ni_Duplicate_A

ssay_Ni Units ResultNo. Pairs: 12 12 Pearson CC: 1.00Minimum: 0.00 0.00 ppm Spearman CC: 1.00Maximum: 344.00 369.00 ppm Mean HARD: 4.34Mean: 90.92 94.17 ppm Median HARD: 2.30Median 47.00 42.00 ppmStd. Deviation: 103.84 110.59 ppm Mean HRD: -1.48Coefficient of Variation: 1.14 1.17 Median HRD -0.79

05

101520

1 10 10 0 10 00

HA

RD

(%

)


Mean vs. HARD Plot(Trench Pulp Duplicate Ni ME-MS81 Analysis


0

50

10 0

0 10 20 30 40 50 60 70 80 90 10 0H

AR

D (

%)

Rank (%)

Rank HARD Plot(Trench Pulp Duplicate Ni ME-MS81 Analysis

Precision: 10%


0

20

40

60

-1.0 0.0 1.0Fre

qu

en

cy (

%)

HRD (/100)

HRD Histogram(Trench Pulp Duplicate Ni ME-MS81 Analysis


-20-10

01020

1 10 10 0 10 00

HR

D (

%)


Mean vs. HRD Plot(Trench Pulp Duplicate Ni ME-MS81 Analysis


0.1

1

10

10 0

10 00

1 10 10 0 10 00

Ab

solu

te D

iffe

ren

ce (

ppm

)


T & H Precision Plot (Assay Pairs)(Trench Pulp Duplicate Ni ME-MS81 Analysis

10% 20% 30% 50%

1

10

10 0

10 10 0Media

n A

D (

ppm

)


T & H Precision Plot (Grouped Pairs)(Trench Pulp Duplicate Ni ME-MS81 Analysis

10% 20% 30% 50%

-20 0

0

20 0

40 0

0 10 0 20 0 30 0 40 0

Du

plica

te_

Ass

ay_

Ni

(ppm

)

Assay_Ni_ (ppm)

Correlation Plot(Trench Pulp Duplicate Ni ME-MS81 Analysis


-20 0

0

20 0

40 0

0 10 0 20 0 30 0 40 0

Du

plica

te_

Ass

ay_

Ni

(ppm

)

Assay_Ni_ (ppm)

QQ Plot(Trench Pulp Duplicate Ni ME-MS81 Analysis

Ref. Line y = 1.06x -2.56

Berlin Project Duplicate Statistics Feb 12(Trench Pulp Duplicate Ni ME-MS81 Analysis )




Assay_Ni_Duplicate_A

ssay_Ni Units ResultNo. Pairs: 10 10 Pearson CC: 1.00Minimum: 2.10 2.30 ppm Spearman CC: 1.00Maximum: 183.50 186.00 ppm Mean HARD: 1.91Mean: 87.40 87.95 ppm Median HARD: 1.42Median 95.70 91.60 ppmStd. Deviation: 73.41 74.61 ppm Mean HRD: -0.11Coefficient of Variation: 0.84 0.85 Median HRD -0.76

0

5

10

1 10 10 0 10 00

HA

RD

(%

)


Mean vs. HARD Plot(Lab Pulp Duplicate Ni ME-MS81 Analysis)


0

50

10 0

0 10 20 30 40 50 60 70 80 90 10 0H

AR

D (

%)

Rank (%)

Rank HARD Plot(Lab Pulp Duplicate Ni ME-MS81 Analysis)

Precision: 10%

020406080

-1.0 0.0 1.0Fre

qu

en

cy (

%)

HRD (/100)

HRD Histogram(Lab Pulp Duplicate Ni ME-MS81 Analysis)


-10-505

10

1 10 10 0 10 00

HR

D (

%)


Mean vs. HRD Plot(Lab Pulp Duplicate Ni ME-MS81 Analysis)


0.1

1

10

10 0

1 10 10 0 10 00

Abso

lute

Dif

fere

nce

(pp

m)


T & H Precision Plot (Assay Pairs)(Lab Pulp Duplicate Ni ME-MS81 Analysis)

10% 20% 30% 50%

T & H Precision Plot (Grouped Pairs)(Lab Pulp Duplicate Ni ME-MS81 Analysis)


0

50

10 0

15 0

20 0

0 20 40 60 80 10 0 12 0 14 0 16 0 18 0 20 0

Du

plica

te_

Ass

ay_

Ni (p

pm

)

Assay_Ni_ (ppm)

Correlation Plot(Lab Pulp Duplicate Ni ME-MS81 Analysis)


0

50

10 0

15 0

20 0

0 20 40 60 80 10 0 12 0 14 0 16 0 18 0 20 0

Du

plica

te_

Ass

ay_

Ni (p

pm

)

Assay_Ni_ (ppm)

QQ Plot(Lab Pulp Duplicate Ni ME-MS81 Analysis)

Ref. Line y = 1.02x -0.82

Berlin Project Duplicate Statistics Feb 12(Lab Pulp Duplicate Ni ME-MS81 Analysis)




Assay_Nd_Duplicate_A

ssay_Nd Units ResultNo. Pairs: 12 12 Pearson CC: 1.00Minimum: 9.10 8.90 ppm Spearman CC: 1.00Maximum: 195.50 196.50 ppm Mean HARD: 0.98Mean: 76.83 77.26 ppm Median HARD: 0.85Median 47.60 49.10 ppmStd. Deviation: 65.08 64.60 ppm Mean HRD: -0.54Coefficient of Variation: 0.85 0.84 Median HRD -0.32

0

5

10

1 10 10 0 10 00

HA

RD

(%

)


Mean vs. HARD Plot(Trench Pulp Duplicate Nd ME-MS81 Analysis


0

50

10 0

0 10 20 30 40 50 60 70 80 90 10 0

HA

RD

(%

)Rank (%)

Rank HARD Plot(Trench Pulp Duplicate Nd ME-MS81 Analysis

Precision: 10%

0

50

10 0

-1.0 0.0 1.0Fre

qu

en

cy (

%)

HRD (/100)

HRD Histogram(Trench Pulp Duplicate Nd ME-MS81 Analysis


-10-505

10

1 10 10 0 10 00

HR

D (

%)


Mean vs. HRD Plot(Trench Pulp Duplicate Nd ME-MS81 Analysis


0.1

1

10

10 0

1 10 10 0 10 00

Ab

solu

te D

iffe

ren

ce (

pp

m)


T & H Precision Plot (Assay Pairs)(Trench Pulp Duplicate Nd ME-MS81 Analysis

10% 20% 30% 50%

0.1

1

10

10 0

10 10 0Med

ian

AD

(p

pm

)


T & H Precision Plot (Grouped Pairs)(Trench Pulp Duplicate Nd ME-MS81 Analysis

10% 20% 30% 50%

050

10 015 020 0

0 20 40 60 80 10 0 12 0 14 0 16 0 18 0 20 0

Du

plica

te_

Ass

ay_

Nd

(pp

m)

Assay_Nd_ (ppm)

Correlation Plot(Trench Pulp Duplicate Nd ME-MS81 Analysis


0

50

10 0

15 0

20 0

0 20 40 60 80 10 0 12 0 14 0 16 0 18 0 20 0

Du

plica

te_

Ass

ay_

Nd

(pp

m)

Assay_Nd_ (ppm)

QQ Plot(Trench Pulp Duplicate Nd ME-MS81 Analysis

Ref. Line y = 0.99x + 1.02

Berlin Project Duplicate Statistics Feb 12(Trench Pulp Duplicate Nd ME-MS81 Analysis )


Appendix B Qualified Persons Certificates

Appendix B – Qualified Persons Certificates Page: 1

Certificate of Qualified Person

Coffey Mining

As an author of the report entitled “Berlin Project, Colombia, National Instrument 43.101 Report” dated

2 March 20121, on the Berlin Project property of U3O8 Corporation (the “Study”), I hereby state:

1. My name is Neil Andrew Inwood and I am a Principal Resource Geologist with the firm of Coffey

Mining Pty. Ltd. of 1162 Hay Street, West Perth, WA, 6005, Australia.

2. I am a practising geologist and a Fellow of the AusIMM (210871).

3. I am a graduate of Curtin University of Technology in Western Australia with a BSc in Geology in

1993 and a PGrad Dip in Hydro-Geology in 1994. In 2007 I graduated from the University of

Western Australia with a MSc in Geology, and from Edith Cowan University with a Post Graduate

Certificate in Geostatistics.

4. I have practiced my profession continuously since 1994.

5. I am a “qualified person” as that term is defined in National Instrument 43-101 (Standards of

Disclosure for Mineral Projects) (the “Instrument”).

6. I visited the Berlin Project property and surrounding areas for 2 days in June 2011. I have

performed consulting services and reviewed files and data supplied by U3O8 Corporation between

June 2011 and March 2012.

7. I contributed to and am responsible for all sections of the Study, apart from Section 13 and the

associated text in the summary, conclusions and recommendations.

8. As of the effective date of the Study, to the best of my knowledge, information and belief, the parts

of the Study for which I am responsible contain all scientific and technical information that is

required to be disclosed to make the Study not misleading.

9. I am independent of U3O8 Corporation pursuant to section 1.4 of the Instrument.

10. I have read the National Instrument and Form 43-101F1 (the “Form”) and the Study has been

prepared in compliance with the Instrument and the Form.

11. I do not have nor do I expect to receive a direct or indirect interest in the Berlin Project property of

U3O8 Corporation, and I do not beneficially own, directly or indirectly, any securities of U3O8

Corporation or any associate or affiliate of such company.

Dated at Perth, Western Australia, on 2nd March, 2012.

[signed]

Neil Inwood Principle Resource Geologist BSc (Geol), MSc Geology), Post Grad Cert Geostatistics



J.R. Goode and Associates

As an author of parts of the report entitled “Berlin Project, Colombia, National Instrument 43.101 Report”

dated 2 March 20121, on the Berlin Project property of U3O8 Corporation (the “Study”), I hereby state:

1. My name is John Richard Goode and I am a Metallurgical Consultant with J.R. Goode and

Associates of Suite 1010, 65 Spring Garden Avenue, Toronto, Ontario, Canada, M2N 6H9.

2. I am a practising metallurgist and a Professional Engineer (Ontario), a member of the CIMM, and a

Fellow of the AusIMM (113128).

3. I am a graduate of the Royal School of Mines, London University with a BSc (Chemical

Engineering in Metallurgy) granted in 1963.




6. I have not visited the Berlin Project property. I have performed consulting services and reviewed

files and data supplied by U3O8 Corporation SGS Minerals Services in Lakefield, Ontario, between

August 2010 and the present time.

7. I contributed to and am responsible for parts of Section 13 of the Study.





10. I have read the National Instrument and Form 43-101F1 (the “Form”) and the parts of the Study for

which I am responsible have been prepared in compliance with the Instrument and the Form.




Dated at Toronto, Ontario, Canada on 2nd March, 2012.

[signed]

G.R. Goode Metallurgical Consultant P.Eng, MCIMM, FAusIMM



As an author of parts of the report entitled “Berlin Project, Colombia, National Instrument 43.101 Report”

dated 2 March 20121, on the Berlin Project property of U3O8 Corporation (the “Study”), I hereby state:

1. My name is Paul Charles Miller and I am a Consulting Metallurgist and Managing Director with the

firm of Sulphide Resource Processing Pty Ltd of 31 Mabena Place, Ocean Reef’ WA 6027, Australia.

2. I am a practising Metallurgist and a Chartered Engineer (U.K) and a Member of the Institute of

Mining and Metallurgy London England.

3. I am a graduate of Birmingham University England with a BSc (Hons) in Minerals Processing in

1978 and an MSC in Biochemical Engineering in 1979 and a Ph.D in Chemical Engineering in

1982. In 1990 I graduated from the University of Cape Town South Africa with an MBA.




6. I have not visited the Berlin Project property. I have performed consulting services and reviewed

files and data supplied by U3O8 Corporation SGS Minerals Services in Lakefield Ontario and SGS

Oretest Minerals Services in Perth Australia between June 2011 and March 2012.

7. I contributed to and am responsible for parts of Section 13 Section 13 of the study.





10. I have read the National Instrument and Form 43-101F1 (the “Form”) and the Parts of the Study for

which I am responsible for have been prepared in compliance with the Instrument and the Form.




Dated at Toronto, Ontario, Canada on 2nd March, 2012.

[signed]

Paul C Miller Managing Director (Consulting Metallurgist) BSc (Hons); MSC; Ph.D (Chem Eng); MBA; C.Eng; MIMM

national instrument ni 43-101 report u3o8 corp. · 1 berlin project, colombia national instrument...

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