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
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 Supervising Principal Neil Inwood Ingvar Kirchner
Berlin Project, Colombia – MINEWPER00790AC National Instrument NI 43-101 Report – 2 March 2012 Coffey Mining Pty Ltd
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] BSc (Geol.), PGradDip (Geol), MSc, FAusIMM Neil Inwood Principle Resource Geologist Coffey Mining Pty Ltd.
Signed on the 2 March 2012
[signed] B.Sc., P.Eng., FAusIMM, MCIMM Mr John Goode Consulting Metallurgist J.R. Goode and Associates
Signed on the 2 March 2012
[signed] Ph.D, MIMM, Chart. Eng. Dr Paul Miller Managing Director Sulphide Resource Processing Pty Ltd
Signed on the 2 March 2012
Berlin Project, Colombia – MINEWPER00790AC National Instrument NI 43-101 Report – 2 March 2012 Coffey Mining Pty Ltd
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
Berlin Project, Colombia – MINEWPER00790AC National Instrument NI 43-101 Report – 2 March 2012 Coffey Mining Pty Ltd
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
Berlin Project, Colombia – MINEWPER00790AC National Instrument NI 43-101 Report – 2 March 2012 Coffey Mining Pty Ltd
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
Berlin Project, Colombia – MINEWPER00790AC National Instrument NI 43-101 Report – 2 March 2012 Coffey Mining Pty Ltd
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
Berlin Project, Colombia – MINEWPER00790AC National Instrument NI 43-101 Report – 2 March 2012 Coffey Mining Pty Ltd
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
Berlin Project, Colombia – MINEWPER00790AC National Instrument NI 43-101 Report – 2 March 2012 Coffey Mining Pty Ltd
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 National Instrument NI 43-101 Report – 2 March 2012 Coffey Mining Pty Ltd
2 SUMMARY0B
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 Contained U O P O V O Y O Mo Ni Ag Re Nd Cutoff Mt 3 8 2 5 2 5 2 3 % U3O8 U3O8 (%) (%) (ppm) (ppm) (%) (ppm) (ppm) (ppm) (% U3O8) (Mkg) (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
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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.
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3 INTRODUCTION1B
3.1 Scope27B 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 Principal28B 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.
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3.3 Participants29B
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 Site30B 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 Qualifications31B 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.
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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 Independence32B
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 Abbreviations3B
A listing of abbreviations used in this report is provided in Table 2.7_1.
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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
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4 RELIANCE2B 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).
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5 PROPERTY3B DESCRIPTION AND LOCATION
5.1 Regulatory34B Framework for Mineral Properties in Colombia
5.1.1 Overview85B
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 Legal86B 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.
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5.1.3 Repatriation87B 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 Canada-Colombia8B Free Trade Agreement
Colombia and Canada signed a free trade agreement in June 2010.
5.1.5 Mineral89B 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.
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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 Other90B 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.
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5.1.7 Concession91B 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
Fees per Hectare (“Ha”) Concession Size 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 Royalties35B
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
NSR Commodity (%) 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 Area36B 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.
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5.4 Location37B 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)
Corner Gauss- Kruger UTM Concession Point Northing Easting Northing Easting 1 1111460 902000 619371 504207 2 1111920 900210 619828 502417 3 1110000 899180 617907 501391 4 1110000 897000 617904 499212 755-17 5 1115000 897000 622901 499204 6 1113000 902000 620910 504205 7 1115000 902070 622909 504272 8 1113000 902070 620910 504275 1 1107000 897830 615220 499671 2 1107000 897000 615219 498842 756-17 3 1110000 897000 618217 498837 4 1110000 899230 618221 501066 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 664-17 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 1 1115310 899109 623215 501312 2 1115310 900001 623216 502203 IFM 08221X 3 1115310 901266 623218 503467 4 1115000 901266 622908 503468 5 1115000 899110 622905 501313 1 1115310 901000 623218 503202 2 1115310 895513 623209 497718 3 1119500 895513 627397 497711 4 1120000 896500 627898 498697 736-17 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.
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5.5 Location38B 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 Details39B Pertaining to the Concessions
5.6.1 Concession92B 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.
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. 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.
5.6.2 Concession93B 2 (File No 756-17)
Legal Status
. Concession Contract:
Executed on 23rd 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:
459 Hectares . Location:
Municipality of Samana, Department of Caldas - Berlin area.
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Commodity
. Uranium and radioactive elements;
. Vanadium, molybdenum and phosphate;
. Any other element of economic value that can be extracted from the concession.
. Current Concession Owner:
Gaia Energy (Colombia) Ltd
. Annual Surface Fee:
USD4,700
. Environmental License:
Not applicable during the exploration stage.
5.6.3 Concession94B 3 (File No 664-17)
Legal Status
. Concession Contract:
Executed on 23rd October 2007;
Registered on 7th December 2007.
Original expiry: 7th December 2010.
. Extension Application:
Application has been made for two-year extension for the concession - expiry: 7th December 2012.
. Mineral concession phase:
Exploration
. Size:
7,304.9 Hectares
. Location:
Municipality of Samana, Department of Caldas – Berlin area. Commodity
. Gold;
. Uranium and radioactive elements;
. Vanadium, molybdenum and phosphate;
. Any other mineral that can be economically exploited in the concession area.
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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
. Environmental License:
Not applicable during the exploration stage.
5.6.4 Concession95B 4 (File No IFM 08221X)
Legal Status
. Concession Contract:
Executed on 12th December 2007;
Registered on 21st December 2007;
Original expiry: 21st December 2010.
. Extension Application:
Application has been made for the two-year extension for the concession – expiry: 21st December 2012.
. Mineral concession phase:
Exploration
. Size:
74.5 Hectares
. Location:
Municipality of Samana, Department of Caldas – Berlin area.
Commodity
. Gold, Silver, copper, zinc, platinum, molybdenum;
. Uranium and radioactive elements;
. Vanadium and phosphate;
. Any other mineral that can be economically exploited from the concession area.
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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
. Environmental License:
Not applicable during the exploration stage.
5.6.5 Concession96B 5 (File No 736-17)
Legal Status
. Concession Contract:
Executed on 23rd October 2007;
Registered on 9th November 2007.
Original expiry: 9th November, 2010.
. Extension Application:
Application has been made for two-year extension for the concession - expiry: 9th November 2012.
. Mineral concession phase:
Exploration
. Size:
2,704 Hectares
. Location:
Municipality of Samana, Department of Caldas - Berlin area. Commodity
. Gold, silver, zinc, platinum, molybdenum
. Uranium and radioactive elements;
. Vanadium and phosphate;
. Any other mineral that can be economically exploited in the concession area.
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. Current Concession Owner:
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
. Environmental License:
Not applicable during the exploration stage.
5.7 Related40B 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 Comments41B
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.
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6 ACCESSIBILITY,4B CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY
6.1 Topography,42B 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
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Figure 5.1_2 Photo showing the Typical Landscape of the Berlin Project
6.2 Access43B 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).
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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 Climate4B
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.
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7 HISTORY5B
7.1 Prior45B 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 Prior46B 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.
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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.
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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.
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Table 6.2_1 Berlin Project Rock-Chip Sample Assay Results from Radioactive Unit, Southern Berlin Project Area
Coordinate Sample Height Radiometry Thickness U3O8 P2O5 Mo V2O5 No. UTM N UTM E m.a.s.l. cps (SPP2) (m) % % % % 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
Location - Coordinates U O Grade Sample Thickness Sample No 3 8 UTM North UTM East (%) (metres) 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)
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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
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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.
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7.3 Historic47B Resource Estimates
7.3.1 Historic97B 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 Metallurgical48B 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.
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8 GEOLOGICAL6B SETTING AND MINERALISATION
8.1 Regional49B 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 Lithology98B 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).
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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)
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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.
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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.
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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 Structure9B
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.
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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 Geology50B of the Berlin Project
8.2.1 Lithology10B & 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.
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Figure 7.2.1_1 Geology of the Berlin Project
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Figure 7.2.1_2 Generalised Stratigraphic Column of the Berlin Project
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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.
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Figure 7.2.1_4 Drillhole DDB-013 Sedimentary Sequence
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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.
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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 Structure10B
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.
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Figure 7.2.2_1 West-East Cross Sections (1 to 4) Through Berlin Syncline and Plan View
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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
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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 Mineralisation51B
8.3.1 Method102B 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 Description103B 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.
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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 Composition104B 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.
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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).
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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).
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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
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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).
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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
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8.3.4 Nature105B 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.
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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).
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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).
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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)
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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.
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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
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8.3.5 Alaskite106B
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.
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8.3.6 Summary107B 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.
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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.
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9 DEPOSIT7B TYPES
9.1 Historic52B 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 Analogous53B 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.
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10 EXPLORATION8B
10.1 Responsibility54B 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 Approach5B
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 Trenching56B
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 Conclusion57B
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.
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Table 9.3_1 Berlin Project Summary of Trench Assay Results
Estimated True U O U O V O P O Mo Re Ag Ni Zn Y O Nd2O3 Trench ID 3 8 3 8 2 5 2 5 2 3 Thickness (%) (lb/st) (%) (%) (ppm) (ppm) (ppm) (ppm) (ppm) (ppm) (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.
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Figure 9.3_1 Geological Map of Southern Berlin Syncline Showing Location of Trenches Completed in 2010 and 2011
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Figure 9.3_2 Geological Map of Northern Berlin Syncline Showing Location of Trenches Completed in 2010 and 2011
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11 DRILLING9B
11.1 Drill58B 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.
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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
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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-007 P3 238 DDB-048 P24 213 DDB-008 P3 131 DDB-049 P24 225 DDB-009 P3 151 DDB-050 P39 255 DDB-010 P4 271 DDB-051 P24 275 DDB-011 P4 166 DDB-052 P40 191 DDB-012 P4 179 DDB-053 P24 215 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-019 P6 280 DDB-060 P23 115 DDB-020 P6 323 DDB-061 P23 150 DDB-021 P30 131 DDB-062 P14 113 DDB-022 P30 151 DDB-063 P14 90 DDB-023 P30 304 DDB-064 P14 108 DDB-024 P30 196 DDB-065 P40 265 DDB-025 P28 207 DDB-066 P14 268 DDB-026 P28 192 DDB-067 P21 201 DDB-027 P28 186 DDB-068 P40 462 DDB-028 P28 230 DDB-069 P21 197 DDB-029 P28 305 DDB-070 P21 285 DDB-030 P29 144 DDB-071 P37 325 DDB-031 P11 331 DDB-072 P21 268 DDB-032 P29 180 DDB-073 P37 382 DDB-033 P29 402 DDB-074 P21 247 DDB-034 P11 346 DDB-075 P37 332 DDB-035 P11 355 DDB-076 P13 152 DDB-036 P27 253 DDB-077 P13 297 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
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Table 10.1_2 Berlin Project Summary of Assay Results from 2010-2011 Drilling Program
Drillhole From To Estimated True U O U O V O P O Mo Re Ag Ni Zn Y O Nd O Platform 3 8 3 8 2 3 2 5 2 3 2 3 ID (m) (m) Width (m) % lb/st (%) (%) (ppm) (ppm) (ppm) (ppm) (ppm) (ppm) (ppm) DDB001 109.7 111.2 1.5 0.079 17393 0.68 14.6 294 1.8 5.9 1140 2930 641 154 P1 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 DDB004 Drillhole did not reach target depth P2 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 DDB007 152.4 155.5 3.1 0.115 25316 0.45 8.9 632 7.3 2.2 1958 2922 386 85 P3 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 DDB010 207.9 211.9 4.0 0.103 2.26 0.4 7.6 584 5 3 2660 3215 402 105 P04 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 DDB015 161.9 163.9 2.0 0.136 2.99 0.5 10.0 658 7 3 2773 3905 518 124 P05 DDB016 182.2 186.2 4.0 0.099 2.18 0.4 7.6 552 6 2 2178 2920 378 93 DDB013 330.7 332.7 2.0 0.131 2.89 0.5 9.2 741 10 4 4768 4740 554 138 P05' DDB014 259.9 261.9 2.0 0.142 3.13 0.5 10.3 725 8 3 3515 4670 605 150 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 P06 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 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 P11 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 DDB076 74.7 77.6 3.0 0.118 2.59 0.5 9.3 675 6 4 2928 3387 511 118 P13 DDB077 116.3 130.8 3.0 0.149 3.29 0.6 10.6 739 7 3 3120 3807 579 127 DDB077* 244.7 245.4 Low Recovery 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 P14 DDB063 Mineralised layer faulted out DDB064 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 DDB042 132.8 135.8 3.0 0.112 2.46 0.4 2.3 563 7 3 2249 3269 467 102 P16 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 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 P17 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 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 P19 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 DDB067 152.7 154.1 1.4 0.111 2.44 0.5 10.2 418 7 3 1800 3217 507 99 P21 DDB067 154.2 155.3 1.1 0.058 1.27 0.3 7.3 184 2 2 1075 2140 479 136
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Table 10.1_2 Berlin Project Summary of Assay Results from 2010-2011 Drilling Program
Drillhole From To Estimated True U O U O V O P O Mo Re Ag Ni Zn Y O Nd O Platform 3 8 3 8 2 3 2 5 2 3 2 3 ID (m) (m) Width (m) % lb/st (%) (%) (ppm) (ppm) (ppm) (ppm) (ppm) (ppm) (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 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 P23 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 DDB048* 164.60 167.00 Low Recovery DDB048 167.4 168.0 0.6 0.185 4.08 0.6 13.4 809 12 3 2410 4200 572 121 P24 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 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 P27 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 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 P28 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 DDB30 Mineralised layer faulted out P29 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 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 P30 DDB023 Mineralised layer faulted out DDB024 Mineralised layer faulted out 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 P37 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 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 P40 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
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11.2 Discussion59B 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.
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12 SAMPLE10B PREPARATION, ANALYSES AND SECURITY
12.1 Qualification60B 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 Laboratory61B 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 Sampling62B Procedure
12.3.1 Exploration108B 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.
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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 Drill109B 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.
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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.
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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
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12.4 Sample63B 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 Sample64B 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 ME-MS61U10B
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.
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12.5.2 ME-M611B
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 ME-MS8112B
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 ME-XRF13B
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 AA2414B
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 Comments65B
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.
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13 DATA1B VERIFICATION
13.1 Verification6B 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 Independent67B 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 Analytical68B 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.
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Table 12.2_1 Berlin Project Independent Sampling Results
Drillhole Length U3O8 (ppm) P2O5 (%) V (ppm) Y (ppm) Mo (ppm) Ni (ppm) Ag (ppm) Nd (ppm) Re (ppm) Sample ID From To ID (m) Original Check Original Check Original Check Original Check Original Check Original Check Original Check Original Check Original Check 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 DDB13 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 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 DDB20 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 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 DDB17 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
Berlin Project, Colombia – MINEWPER00790AC Page 77 National Instrument NI 43-101 Report – 2 March 2012 Coffey Mining Pty Ltd
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 U3O815B 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
U (ppm) V (ppm) Mo (ppm) P (%) Y (ppm) Standard Exp Std Exp Std Exp Std Exp Std Exp Std Value Dev Value Dev Value Dev Value Dev Value 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
Berlin Project, Colombia – MINEWPER00790AC Page 78 National Instrument NI 43-101 Report – 2 March 2012 Coffey Mining Pty Ltd
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%.
Berlin Project, Colombia – MINEWPER00790AC Page 79 National Instrument NI 43-101 Report – 2 March 2012 Coffey Mining Pty Ltd
Figure 12.3.1_1 Berlin Project Uranium Standard Control Plots
Standard Control Plot Standard Control Plot (Standard ST1000020 ME-MS61 Analysis ) (Standard ST1000020 ME-MS61U Analysis)
2200 2200
2100 2100 2000
2000 1900
1800 1900 U_Result (ppm) U_Result U_Result (ppm) U_Result 1700 1800 1600
1500 02-Feb-2011 09-Apr-2011 01-Jul-2011 18-Aug-2011
170015-Nov-2010 18-Nov-2010 18-Nov-2010 02-Dec-2010 02-Dec-2010 06-Dec-2010 16-Dec-2010 10-Jan-2011
DateReported DateReported
U_Result Expected Value = 1,910.20 EV Range (1,719.18 to 2,101.22) Mean of U_Result = 1,848.89 U_Result Expected Value = 1,910.20 EV Range (1,719.18 to 2,101.22) Mean of U_Result = 1,802.44
Standard Control Plot Standard Control Plot (Standard ST1000045 ME-MS61U Analysis ) (Standard AMIS-055 ME-MS61U Analysis )
1600 3700 3600 1500 3500 3400 1400 3300 3200 1300 U_Result (ppm)
U_Result (ppm) U_Result 3100
1200 3000 2900
1100 17-May-2011 01-Jul-2011 30-Jul-2011 26-Sep-2011
280020-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
DateReported DateReported
U_Result Expected Value = 1,353.00 EV Range (1,217.70 to 1,488.30) Mean of U_Result = 1,259.76 U_Result Expected Value = 3,206.00 EV Range (2,885.40 to 3,526.60) Mean of U_Result = 3,360.71
Berlin Project, Colombia – MINEWPER00790AC Page 80 National Instrument NI 43-101 Report – 2 March 2012 Coffey Mining Pty Ltd
Figure 12.3.1_2 Berlin Project Vanadium Standard Control Plots
Standard Control Plot Standard Control Plot (Standard ST1000020 V ME- MS61 Analysis) (Standard ST1000045 V ME- MS61U Analysis )
3800 4000 3700 3900 3800 3600 3700 3500 3600 3400 3500
3300 3400 V_Result (ppm) V_Result (ppm) 3300 3200 3200 3100 3100
3000 17-May-2011 01-Jul-2011 30-Jul-2011 26-Sep-2011
300015-Nov-2010 18-Nov-2010 18-Nov-2010 02-Dec-2010 02-Dec-2010 06-Dec-2010 16-Dec-2010 10-Jan-2011
DateReported DateReported
V_Result Expected Value = 3,414.00 EV Range (3,072.60 to 3,755.40) Mean of V_Result = 3,274.44 V_Result Expected Value = 3,584.00 EV Range (3,225.60 to 3,942.40) Mean of V_Result = 3,348.54
Standard Control Plot Standard Control Plot (Standard ST1000020 V ME- MS61U Analysis ) (Standard ST1000020 V ME- MS81Analysis )
3800 4000 3900 3700 3800 3600 3700 3500 3600 3500 3400 3400 3300 3300 V_Result (ppm) V_Result (ppm) 3200 3200 3100 3100 3000
3000 02-Feb-2011 09-Apr-2011 01-Jul-2011 18-Aug-2011
2900 11-Jan-2011 28-Feb-2011 19-Apr-2011
DateReported DateReported
V_Result Expected Value = 3,414.00 EV Range (3,072.60 to 3,755.40) Me an of V_Res ult = 3,203.86 V_Result Expected Value = 3,414.00 EV Range (3,072.60 to 3,755.40) Mean of V_Result = 3,393.94
Berlin Project, Colombia – MINEWPER00790AC Page 81 National Instrument NI 43-101 Report – 2 March 2012 Coffey Mining Pty Ltd
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 3,072.60 to 3,072.60 to 3,072.60 to 3,225.60 to 3,225.60 to Expected Value Range 3,755.40 3,755.40 3,755.40 3,942.40 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%
Berlin Project, Colombia – MINEWPER00790AC Page 82 National Instrument NI 43-101 Report – 2 March 2012 Coffey Mining Pty Ltd
Figure 12.3.1_3 Berlin Project Molybdenum Standard Control Plots
Standard Control Plot Standard Control Plot (Standard ST1000020 Mo ME-MS61U Analysis) (Standard ST1000020 Mo ME-MS81 Analysis)
80 80 70 70 60 60 50 Mo_Result (ppm) Mo_Result Mo_Result (ppm) Mo_Result
40 17-May-2011 01-Jul-2011 30-Jul-2011 26-Sep-2011
50 02-Feb-2011 09-Apr-2011 01-Jul-2011 18-Aug-2011
DateReported DateReported
Mo_Result Expected Value = 64.18 Mo_Result Expected Value = 64.18 EV Range (57.76 to 70.60) Mean of Mo_Result = 57.85 EV Range (57.76 to 70.60) Mean of Mo_Result = 65.59 Standard Control Plot (Standard ST1000020 Mo ME-MS61 Analysis )
75 70 65 60 Mo_Result (ppm) Mo_Result
5515-Nov-2010 18-Nov-2010 18-Nov-2010 02-Dec-2010 02-Dec-2010 06-Dec-2010 16-Dec-2010 10-Jan-2011
DateReported
Mo_Result Expected Value = 64.18 EV Range (57.76 to 70.60) Mean of Mo_Result = 66.71
Berlin Project, Colombia – MINEWPER00790AC Page 83 National Instrument NI 43-101 Report – 2 March 2012 Coffey Mining Pty Ltd
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%
Berlin Project, Colombia – MINEWPER00790AC Page 84 National Instrument NI 43-101 Report – 2 March 2012 Coffey Mining Pty Ltd
Figure 12.3.1_4 Berlin Project Phosphorus Standard Control Plots
Standard Control Plot Standard Control Plot (Standard ST1000020 P ME-MS81 Analysis) (Standard ST1000020 P ME-XRF12 Analysis )
1.9 1.9 1.8 1.8 1.7 1.7 1.6 P_Result (%) P_Result
P_Result (%) 1.6
1.5 27-Jan-2011 12-Mar-2011 08-Jun-2011
1.5 11-Jan-2011 28-Feb-2011 19-Apr-2011
DateReported DateReported
P_Result Expected Value = 1.76 P_Result Expected Value = 1.76 EV Range (1.58 to 1.94) Mean of P_Result = 1.87 EV Range (1.58 to 1.94) Mean of P_Result = 1.87 Standard Control Plot Standard Control Plot (Standard ST1000045 P ME-MS81 Analysis ) (Standard ST1000045 P ME-XRF12 Analysis )
2.6 2.6 2.5 2.5 2.4 2.3 2.4
P_Result (%) P_Result 2.2
P_Result (%)2.3
2.1 19-Apr-2011 19-Apr-2011 07-May-2011 16-May-2011 17-May-2011 08-Jun-2011 08-Jul-2011 18-Jul-2011 20-Jul-2011
2.2
2.123-Feb-2011 19-Apr-2011 19-Apr-2011 19-Apr-2011 21-Apr-2011 17-May-2011 26-May-2011 18-Jul-2011
DateReported
DateReported P_Result Expected Value = 2.44 EV Range (2.20 to 2.68) Mean of P_Result = 2.49 P_Result Expected Value = 2.44 EV Range (2.20 to 2.68) Mean of P_Result = 2.49
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Figure 12.3.1_5 Berlin Project Phosphorus Standard Control Plot Standard Control Plot (Standard AMIS-055 P ME-XRF12 Analysis)
10.4 10.2 10.0 9.8 9.6 9.4 9.2 P_Result (%) 9.0 8.8 8.6
8.418-Aug-2011 19-Aug-2011 26-Aug-2011 05-Sep-2011 19-Sep-2011 19-Sep-2011 01-Oct-2011 01-Oct-2011
DateReported
P_Result Expected Value = 9.36 EV Range (8.42 to 10.30) Mean of P_Result = 9.36
Figure 12.3.1_6 Berlin Project Yttrium Standard Control Plots Standard Control Plot (AMIS-055 Y ME-MS81 Analysis )
60 58 56 54 52 50 Y_Result (ppm)
4820-Jul-2011 19-Aug-2011 26-Aug-2011 05-Sep-2011 19-Sep-2011 01-Oct-2011 01-Oct-2011
DateReported
Y_Result Expected Value = 53.60 EV Range (48.24 to 58.96) Mean of Y_Result = 55.41 Standard Control Plot (AMIS-055 Y ME-MS61U Analysis )
62 60 58 56 54 52 50 Y_Result (ppm)
4820-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
DateReported
Y_Result Expected Value = 53.60 EV Range (48.24 to 58.96) Mean of Y_Result = 56.29
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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 Drillhole16B 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 Analysis No of Mean Median Within RANK HARD Limit Comparative Means (ppm) Element Type Method Pairs % HARD % HARD (10%/20%) Original/Duplicate ME-MS81 8 2.04 1.47 100/100 318/332 U ME-MS61U 15 4.35 2.4 93/93 248/248 ME-MS81 8 2.25 0.77 100/100 1,899/1,939 V ME-MS61U 15 2.56 1.55 100/100 1,179/1,171 ME-MS81 8 3.49 2.6 88/100 371/171 Mo ME-MS61U 15 5.98 3.08 80/100 199/199 Drillhole ME-MS81 8 2.78 1.5 100/100 1118/119 Y 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 15 2.54 2.2 100/100 664/659 P 15 5.5 1.15 87/93 5,287/5,117 ME-MS61U 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.
Berlin Project, Colombia – MINEWPER00790AC Page: 87 National Instrument NI 43-101 Report – 2 March 2012 Coffey Mining Pty Ltd
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 Analysis No of Mean Median Within RANK HARD Limit Comparative Means (ppm) Element Type Method Pairs % HARD % HARD (10%/20%) Original/Duplicate U 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 Trenches ME-MS81 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 Analysis No of Mean Median Within RANK HARD Limit Comparative Means (ppm) Element Type Method Pairs % HARD % HARD (10%/20%) Original/Duplicate U 7 1.3 1.33 100/100 378/378 V 10 1.99 1.19 100/100 2,889/2,876 Lab P ME-MS81 10 2.15 0.64 100/100 783/764 Duplicates Y 10 0.98 0.87 100/100 342/343 Ni 10 1.91 1.42 100/100 87/88
Berlin Project, Colombia – MINEWPER00790AC Page: 88 National Instrument NI 43-101 Report – 2 March 2012 Coffey Mining Pty Ltd
13.4 Comments69B
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.
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14 MINERAL12B PROCESSING AND METALLURGICAL TESTING
14.1 Approach70B
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 Laboratories71B 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.
Berlin Project, Colombia – MINEWPER00790AC Page: 90 National Instrument NI 43-101 Report – 2 March 2012 Coffey Mining Pty Ltd
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 Nature72B 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.
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Table 13.3_1 Berlin Project Details of Drillhole Intercepts Used in Metallurgical Testwork to Date
From To Mass Composite Sample Lab Material Facies Drillhole Number Sample Type (m) (m) (kg) Weathered, saprolite-like DDB-001 109.7 111.2 3 Colombian Comp Sandstone & Siltstone ¼ core material DDB-002 77.7 85.1 5 DDB5 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 SGS Lakefield DDB-10 207.89 211.89 Rock from drill core DDB-011 123.2 125.94 DDB-012 135.46 138.2 DDB10-15 Limestone 10# & ¼ core 42 DDB-013 328.68 332.68 DDB-014 257.98 261.9 DDB-015 161.14 163.86 DDB-018 226.27 229.77 DDB-019 225.78 227.78 UC Comp SGS OreTest Rock from drill core Limestone DDB-020 294.63 298.53 10# & ¼ core 11 DDB-025 155.44 159.09 DDB-026 158.42 161.32 DDB-016 181.45 186.95 DDB-017 236.86 240.94 DDB-018 226.27 229.77 BER-160611 DDB-019 225.22 228.35 ¼ core 39 DDB-020 294.63 298.53 DDB-025 155.44 159.09 DDB-026 158.42 162.22 DDB-027 160.32 163.16 ANSTO Rock from drill core Limestone DDB-028 180.95 183.55 DDB-029 253.79 258.93 DDB-031 295.13 297.66 DDB-033 370.45 373.15 DDB27-38 or ANSTO 2 ¼ or ½ core 49 DDB-034 296.35 300.3 111.25 115.2 DDB-036 222.93 225.0 94.69 97.53 DDB-038 287.91 290.14
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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 Head73B 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.
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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 Composite Sample Names Heads Element Units DDB7 DDB5 Berlin DDB10-15 ANSTO 1 or UC ANSTO 2 or Colombian Comp Comp Comp Comp BER-160611 Comp DDB27-38 Comp
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
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14.5 Metallurgical74B Testwork
14.5.1 Baseline17B 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 Acetic18B 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).
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Table 13.5.1_1 Berlin Project Composite Sample BER-16061 - Summary of Alkaline Leach Conditions and Metal Extractions Achieved
Conditions Extractions (%) Na CO NaHCO Oxidant Test No. Time Grind Temp Pulp Density 2 3 3 pH Addn Addn Addn U V P Y Mo Ni Hrs p80µm °C % Solids kg/t kg/t kg/t 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
Acid Metal Recovery (%) Pulp Leach Lab Grind P80 T Average Net Acid Laboratory Sample Density Time Acid Added Test # (µm) (°C) pH Type Consumption U V P Y Mo Ni Re (%) (Hours) (kg/t) (kg/t)
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 SGS 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
BER 1A BER-160611 35 60 2 24 1.5 H2SO4 477 468 76.8 24.6 28.1 43 40 ANSTO 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.
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Table 13.5.2_1 Berlin Project Summary of Acetic Acid Pre-Leach Conditions and Subsequent Strong Acid Tests and Metal Extractions Achieved
Acid Metal Recovery (%) Pulp Lab Test Grind P80 T Average Net Acid Laboratory Sample Density Concentration Acid added # (µm) (°C) pH Type Consumption U V P (%) (g/L) (kg/t) (kg/t) 32 DDB10-15 20 33 5.1 Acetic 385 4.7 1
SGS 55 Test 32 low-carbonate 100 81 25 3.69 H2SO4 50 715 595 95.3 84 74 56 Residue 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
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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 Ferric19B 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).
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Table 13.5.3_1 Berlin Project Summary of Ferric Leach Tests Undertaken by SGS OreTest on Two Composite Samples From Berlin
Ferric Leach (Stage1) Releach / Wash (Stage 2) Elemental Extraction % Maximum Composite Pulp Pulp Grain Size Concentration Temp. Acid Type Temp. Sample # Density Density Uranium Vanadium Phosphate Yttrium Neodymium Zinc Nickel Molybdenum Rhenium (µm) (g/L) (°C) & Conc. (°C) (%) (%) 106 50 10 10%HCL 10 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 UC Comp 65 40 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 42.8 106 50 5 10%HCL 10 97.2 80.2 99.5 95.8 82.8 98.5 62.7 56.8 27.1 ANSTO 2 65 40 106 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 10%H2SO4 10 93.1 56.0 97.9 79.6 48.9 93.7 58.6 44.8 22.8
106 50 5 65 10%H2SO4 10 40 97.3 69.4 99.4 89.8 63.6 97.0 62.7 53.8 11.1 ANSTO 2 106 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
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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.
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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
Maximum Pulp Elemental Extraction % Composite Concentration Temp. Final pH Grain Size Density Sample # (g/L) (°C) (after 48hrs) (µm) (%) Uranium Vanadium Phosphate Yttrium Neodymium Zinc Nickel Molybdenum Rhenium 106 50 10 40 1.9 50.9 2.5 1.8 9.6 nd nd 34.2 7.0 nd UC Comp 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
Ferric Leach (Stage 1) Re-Leach (Stage 2) Combined Composite Pulp Density Time Uranium Extraction Leach Test Acid Type Uranium Stage 1 and 2 Sample Leach Test Number (% solids) (h) (%) Number and Conc. Extraction (%) Extraction (%)
24 79.8 43 10% H2SO4 89.5 97.9
39 10 47 10% H2SO4 31.2 86.7 48 80.7 51 10% HCl 90.8 98.2 DDB10-15 24 82.1 44 10% H2SO4 79.6 96.3
40 5 48 10% H2SO4 62.8 95.5 48 87.8 52 10% HCl 89.6 98.7
24 85.5 45 10% H2SO4 66.8 95.2
41 10 49 10% H2SO4 78.8 94.0 48 71.8 53 10% HCl 76.0 93.2 DDB5 24 70.9 46 10% H2SO4 96.3 98.9
42 5 50 10% H2SO4 70.6 95.7 48 85.4 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
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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):
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. 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.
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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
Kg acid / t Kg Acid / t to Total Overall Fe Fe Fe Lost Initial ferric Kg/t Ferric Feed provided by Fe Kg Acid / t be Consumed kg/t Acid Make- Mass Fe Lixiviant Fe Lixiivant Converted Re-Mobolised (neither Lixiviant Lixiviant % Uranium Test ID Sample Density Acid Target, Conc pptation and Consumed in Reconverting up for Releach Increase Unreacted pptated to Ferrous in Acid Re- mobilised nor Inventory Make-up Extracted (%) Ferrous Acid Releach Ferrous Back + Ferrous Solution Leach as Ferrous) kg/t Required Generation to Ferric Conversion 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% 80.7% 18.6% 450.0 83.7 1070.2 124.6 34.4 159.0 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%
Test 39 DDB10-15 Comp 10 50g/L Fe3+ added as ferric sulphate 84.0% 3.9% 87.4% 8.7% 80.7% 65.3% 450.0 294.0 1070.2 0.0 34.4 34.4 Test 47 Releach Test 39 RES 48h 10 10% Sulphuric for Re-leach -3.6% 22.1% 31.2% %overall wt increase 77.4% %U Total = 86.8%
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% Test 39 DDB10-15 Comp 10 50g/L Fe3+ added as ferric sulphate 84.0% 3.9% 87.4% 8.7% 80.7% 25.3% 450.0 113.9 1070.2 121.4 34.4 155.9 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% 87.8% 0.8% 950.0 7.6 786.5 654.1 64.1 718.2 Test 44 Releach Test 40 RES 24h 10 10% Sulphuric for Re-leach -42.6% 28.1% 79.6% %overall wt increase 15.4% %U Total = 97.5%
Test 40 DDB10-15 Comp 5 50g/L Fe3+ added as ferric sulphate 101.0% 37.4% 28.9% 7.7% 87.8% 8.4% 950.0 79.6 786.5 794.4 64.1 858.5 Test 48 Releach Test 40 RES 48h 10 10% Sulphuric for Re-leach -24.7% 20.5% 62.8% %overall wt increase 51.3% %U Total = 95.5%
Test 40 DDB10-15 Comp 5 50g/L Fe3+ added as ferric sulphate 101.0% 37.4% 28.9% 7.7% 87.8% 2.4% 950.0 22.4 786.5 182.3 64.1 246.5 Test 52 Releach Test 40 RES 48h 10 10% HCl Hydrochloric for Re-leach 61.1% 26.5% 89.6% %overall wt increase 223.8% %U Total = 98.7%
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% 71.8% 1.1% 450.0 4.9 169.4 124.0 73.0 197.1 Test 45 Releach Test 41 RES 24h 10 10% Sulphuric for Re-leach -15.8% 7.0% 66.8% %overall wt increase -15.7% %U Total = 90.6%
Test 41 DDB5 Comp 10 50g/L Fe3+ added as ferric sulphate 0.1% 73.4% 8.1% 18.5% 71.8% 0.9% 450.0 3.9 169.4 194.6 73.0 267.6 Test 49 Releach Test 41 RES 48h 10 10% Sulphuric for Re-leach -6.7% 7.3% 78.8% %overall wt increase -6.6% %U Total = 94.0%
Test 41 DDB5 Comp 10 50g/L Fe3+ added as ferric sulphate 0.1% 73.4% 8.1% 18.5% 71.8% 0.7% 450.0 2.9 169.4 96.0 73.0 169.0 Test 53 Releach Test 41 RES 48h 10 10% HCl Hydrochloric for Re-leach -37.1% 7.5% 76.0% %overall wt increase 0.0% %U Total = 93.2%
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% 85.4% 2.1% 950.0 19.7 739.5 N/A 78.6 N/A Test 46 Releach Test 42 RES 24h 10 10% Sulphuric for Re-leach -40.2% 24.4% 96.3% %overall wt increase -43.5% %U Total = 99.5%
Test 42 DDB5 Comp 5 50g/L Fe3+ added as ferric sulphate -5.5% 64.1% 26.4% 9.4% 85.4% 0.3% 950.0 3.1 739.5 138.6 78.6 217.2 Test 50 Releach Test 42 RES 48h 10 10% Sulphuric for Re-leach -15.2% 26.1% 70.6% %overall wt increase -19.9% %U Total = 95.7%
Test 42 DDB5 Comp 5 50g/L Fe3+ added as ferric sulphate -5.5% 64.1% 26.4% 9.4% 85.4% 0.3% 950.0 3.0 739.5 33.3 78.6 111.9 Test 54 Rleach Test 42 RES 48 h 10 10% HCl Hydrochloric for Re-leach -32.3% 26.1% 71.5% %overall wt increase -36.0% %U Total = 95.8%
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14.5.4 Flotation120B 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.
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. 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 702kg/tonne of feed or approximately 304kg/tonne of ore (Table 13.5.4_2).
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 725kg/tonne of feed or approximately 370kg/tonne of ore (Table 13.5.4_2).
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.
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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
Mass % Distribution Test No Flotation Product Total Organic % U O Calcite Apatite Dolomite Sulphur 3 8 Carbon 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 F1 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 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 F2 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 Slimes 14.0 29.7 10.8 9.8 20.4 Rougher Concentrate 11.8 17.9 11.1 19.9 12.1 SB-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 Slimes 14.0 26.8 11.1 9.6 21.2 Rougher Concentrate 16.9 21.8 16.3 22.0 21.2 SB-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 Slimes 14.0 28.0 11.5 9.5 21.5 Rougher Concentrate 11.6 14.0 9.6 21.7 14.2 SB-3 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 Slimes 14.0 27.5 10.9 9.5 20.4 Rougher Concentrate 31.8 37.0 28.9 41.2 34.2 SB-4 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 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 DDB-F1 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 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 DDB-F2 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 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 DDB-F3 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 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 DDB-F4 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 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 DDB-F5 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 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 DDB-F6 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
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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
Sulphuric Acid % Distribution in Sample Extraction from Leach Feed (kg per tonne of Feed) Sulphuric Acid Consumption Leach Test Sample per Tonne of Ore Mass U O U O V O 3 8 3 8 2 5 Added Consumed (kg/tonne) (%) (%) (%) (%) 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
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Figure 13.5.4_1 Leach Kinetics on Flotation Products from Composite Sample DDB10-15 Berlin flotation products Uranium leach kinetics 100 90 80 70 (%) 60
50 DDBF1 C Prefloat
extracted 40 DDBF1 Con U 30 DDBF2 C Prefloat 20 DDBF2 Con 10 ‐ 0 1020304050607080 Leach time (h)
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
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14.6 Summary75B 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.
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. 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.
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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.
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15 MINERAL13B 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 Resource76B Database and Validation
15.1.1 Database12B
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 Validation12B
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.
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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 Geological7B Interpretation and Modelling
15.2.1 Geological123B 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 Weathering78B 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.
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Figure 14.2.1_1 Modelled Mineralisation Used for 2012 Resource
Showing mineralised outline, drilling and trench locations
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Figure 14.2.1_2 Oblique Section 618000mN with Drilling and U3O8 Values
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15.4 Statistical79B Analysis
15.4.1 Radiometric124B 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 Summary125B 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.
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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 Bulk126B 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).
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Table 14.4.2_1 Berlin Project Summary Mineralised Zone Composite Statistics
U O P O V O Y O Mo Ni Ag Re Nd 3 8 2 5 2 5 2 3 (ppm) (%) (ppm) (ppm) (ppm) (ppm) (ppm) (ppm) (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
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Figure 14.4.2_1 Multi Element Histogram Plots – Mineralised Zone
Histogram Plot Histogram Plot Histogram Plot (Mineralised Zone) (Mineralised Zone) (Mineralised Zone)
6 11 4
10
5 9
3 8 4 7
6 3 2 5 Frequency (%) Frequency (%) Frequency Frequency (%) Frequency 4 2
3 1
1 2
1
0 0 0 0 1000 2000 3000 -1 0 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 1000 2000 3000 4000 5000 6000 7000 8000 9000 10000 11000 resc_u3o8 (g/t) P2O5_pct V2O5_ppm
Histogram Plot Histogram Plot Histogram Plot (Mineralised Zone) (Mineralised Zone) (Mineralised Zone)
6 5 6
5 5 4
4 4
3
3 3 Frequency (%) Frequency Frequency (%) Frequency Frequency (%) Frequency 2
2 2
1 1 1
0 0 0 0 100 200 300 400 500 600 700 800 900 1000 1100 1200 1300 0 100 200 300 400 500 600 700 800 900 1000 1100 1200 1300 1400 1500 1600 0 1000 2000 3000 4000 5000 6000 7000 8000 9000 Y2O3_ppm Mo_ppm Ni_ppm
Histogram Plot Histogram Plot Histogram Plot (Mineralised Zone) (Mineralised Zone) (Mineralised Zone)
28 18 4
26 16 24
22 14 3 20 12 18
16 10 14 2
12 8 Frequency (%) Frequency (%) Frequency (%) 10 6 8 1 6 4
4 2 2
0 0 0 -1012345678910111213141516171819202122232425 0 102030 100 200 300 Ag_ppm Re_ppm Nd_ppm
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Figure 14.4.2_2 Scatter Plots for U3O8 (ppm) Compared to Other Elements
P2O5 (ppm) V2O5 (ppm)
220000 12000
11000 200000
10000 180000
9000 160000 8000 140000 7000 120000 6000 100000
V2O5 (ppm) V2O5 5000 P2O5 (ppm) P2O5
80000 4000
60000 3000
40000 2000
20000 1000
0 0 0 1000 2000 3000 400 0 1000 2000 3000 400 resc_u3o8 (ppm) resc_u3o8 (ppm)
PCC=081SCC=086 Ref Line y = 2 78x + 1 590 46 P.CC= 0.85 S.CC= 0.88 Ref. Line y = 55.40x + 28,483.10
Y2O3 (ppm) Mo (ppm)
4000 4000
3000 3000
2000 2000 Y2O3 (ppm) Mo (ppm)
1000 1000
0 0 1000 2000 3000 4000 0 resc_u3o8 (ppm) 0 1000 2000 3000 4000 resc_u3o8 (ppm)
P.CC= 0.88 S.CC= 0.89 Ref. Line y = 0.38x + 67.84 P.CC= 0.63 S.CC= 0.72 Ref . Line y = 0.30x + 265.90
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Figure 14.4.2_3 Scatter Plots for U3O8 (ppm) Compared to Other Elements Ni (ppm) Ag (ppm)
10000 40
9000
8000 30 7000
6000
5000 20 Ni (ppm) Ag (ppm) Ag 4000
3000 10 2000
1000
0 0 0 1000 2000 3000 4000 0 1000 2000 3000 4000 resc_u3o8 (ppm) resc_u3o8 (ppm)
P.CC= 0.64 S.CC= 0.75 Ref. Line y = 1.56x + 581.59 P.CC= 0.50 S.CC= 0.72 Ref . Line y = 0.00x + 1.16 Nd (ppm) Re (ppm)
400 40
300 30
200 20 Nd (ppm) Re (ppm)
100 10
0 0 0 1000 2000 3000 4000 0 1000 2000 3000 4000 resc_u3o8 (ppm) resc_u3o8 (ppm)
P.CC= 0.77 S.CC= 0.78 Ref . Line y = 0.07x + 19.88 P.CC= 0.69 S.CC= 0.83 Ref. Line y = 0.00x + 1.16
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Table 14.4.3_1 Berlin Project Bulk Density Data Summary Statistics (t/m³) (Diameter Method)
Average U O Min Max Average Region Count 3 8 (ppm) Density Density 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 Variography127B
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.
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Figure 14.4.3_1 Berlin Density Histograms (Diameter Method)
Histogram Plot (Mineralisation)
40
30
20 Frequency (%) Frequency
10
0 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 diameter (t/m3)
Histogram Plot (Limestone)
50
40
30
20 Frequency (%) Frequency
10
0 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 diameter (t/m3)
Histogram Plot (Shale)
30
20 Frequency (%) Frequency 10
0 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 diameter (t/m3)
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Figure 14.4.3_2 Berlin Density Histograms (Diameter Method)
Histogram Plot (Carbonaceous Shale)
50
40
30
20 Frequency (%) Frequency
10
0 1234 diameter (t/m3)
Histogram Plot (Sandstone)
40
30
20 Frequency (%) Frequency
10
0 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 diameter (t/m3)
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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
Semi- Semi- Major Minor Major Minor Region Nugget C1 Major C2 Major Range Range Range Range Range Range Mineralisation 10% 65% 60 60 60 25% 180 180 180
15.5 Block80B 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.
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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 Resource Classification; 1 = Measured, 2 = Indicated, 3 – Inferred, category Integer 0 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 Grade81B Estimation Parameters
15.6.1 U128B 3O8 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.
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Table 14.6.1_1 Berlin Project
U3O8 Sample Search Parameters – Ordinary Kriging
Search Orientation Search Radii Number of Samples Major Semi- Minor Zones Pass Semi- Max / Major Minor Axis Major Axis Min Max Major Hole (m) Axis (m) (m) 1 100 67 67 8 16 4 Mineralisation 2 0 330° 00° 00° 200 133 133 8 16 4 3 400 267 267 3 12 4 Discretisation of 5 x 5 x 3 used
15.6.2 Multi-Element129B 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
Search Orientation Search Radii Number of Samples Major Semi- Minor Zones Pass Semi- Max / Major Minor Axis Major Axis Min Max Major Hole (m) Axis (m) (m) 1 200 67 67 4 12 3 Mineralisation 0 330° 00° 00° 2 500 335 335 2 12 3 Discretisation of 3 x 3 x 3 used
15.7 Bulk82B 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 Resource83B Reporting and Classification
15.8.1 Introduction130B
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.
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15.8.2 Criteria13B 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 Industry standard for diamond drilling. Moderate to High and Sample Preparation Good internal laboratory and external quality control data available for Moderate to High Quality of Assay Data the majority of the chemical assaying. More bulk density data is required. Moderate Verification of Sampling and QAQC analysis is within industry acceptable standards. Moderate to High Assaying The bulk of drill collars surveyed by DGPS and most drillholes have been Location of Sampling Points High downhole surveyed. 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 The interpreted lithological and mineralisation boundaries are considered Geological Interpretation Moderate robust and of good confidence. Estimation and Modelling Estimates based on detailed statistical and geostatistical analysis. Moderate to High Techniques Cutoff Grades 400ppm mineralisation targeted. Moderate Whole block estimates for all mineralised regions completed. No mining Mining Factors or has been undertaken in the deposit. Open pit and underground mining Moderate Assumptions is conceptualised for this project.
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.
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Figure 14.8.2_1 Drillhole Locations, Mineralisation Model and Classification Plan View
Oblique View
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15.8.3 Grade132B 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 Contained U O P O V O Y O Mo Ni Ag Re Nd Cutoff Mt 3 8 2 5 2 5 2 3 % U3O8 U3O8 (%) (%) (ppm) (ppm) (%) (ppm) (ppm) (ppm) (% U3O8) (Mkg) (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 Conclusions84B
. 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.
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16 MINERAL14B RESERVE ESTIMATES
No reserves have been stated for this project.
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17 MINING15B 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.
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18 RECOVERY16B METHODS
Not Applicable.
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19 PROJECT17B 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.
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20 MARKET18B STUDIES AND CONTRACTS
U3O8 Corp. has not currently undertaken any market studies for the project.
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21 ENVIRONMENTAL19B 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.
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22 CAPITAL20B AND OPERATING COSTS
Not applicable
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23 ECONOMIC21B ANALYSIS
Not applicable
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24 ADJACENT2B PROPERTIES
Not applicable.
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25 OTHER23B 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.
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26 INTERPRETATION24B 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.
Berlin Project, Colombia – MINEWPER00790AC Page: 142 National Instrument NI 43-101 Report – 2 March 2012 Coffey Mining Pty Ltd
27 RECOMMENDATIONS25B
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
Estimated Cost Item (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
Berlin Project, Colombia – MINEWPER00790AC Page: 143 National Instrument NI 43-101 Report – 2 March 2012 Coffey Mining Pty Ltd
28 REFERENCES26B
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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.
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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.
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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.
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Vakhrameev, V.A., 1991. Jurassic and Cretaceous floras and climates of the earth, Cambridge University Press, London, 340 p.
Berlin Project, Colombia – MINEWPER00790AC Page: 146 National Instrument NI 43-101 Report – 2 March 2012
Appendix A
QAQC Summary Plots
Berlin Project Uranium Standards Feb 2012
Berlin Project Standards Feb 2012 (Standard ST1000020 ME-MS61 Analysis )
Standard: ST1000020 No of Analyses: 9 Element: U Result Minimum: 1,760.00 Units: ppm Max imum: 1,960.00 Detection Limit: 1 Mean: 1,848.89 Expected Value (EV): 1,910.20 Std Deviation: 66.07 E.V. Range: 1,719.18 to 2,101.22 % in Tolerance 100.00 % % Bias -3.21 % % RSD 3.57 %
Standard Control Plot (Standard ST1000020 ME-MS61 Analysis )
2200 2100 2000 1900 1800 U_Result (ppm) U_Result
170015-Nov-2010 18-Nov-2010 18-Nov-2010 02-Dec-2010 02-Dec-2010 06-Dec-2010 16-Dec-2010 10-Jan-2011
DateReported
U_Result Expected Value = 1,910.20 EV Range (1,719.18 to 2,101.22) M ean of U_Result = 1,848.89
Cumulative Deviation from Assay Mean (Standard ST1000020 ME-MS61 Analysis )
100 50 0 -50 -10 0
-15 015-Nov-2010 18-Nov-2010 18-Nov-2010 02-Dec-2010 02-Dec-2010 06-Dec-2010 16-Dec-2010 10-Jan-2011 Cumulative Sum of U_Result - Mean (ppm) - Mean U_Result of Sum Cumulative
DateReported
U_Result M ean of Cumulative Sum of U_Result - Mean (ppm) = -24. 44 p
Cumulative Deviation from Expected Value (Standard ST1000020 ME-MS61 Analysis )
0 -10 0 -20 0 -30 0 -40 0 -50 0
-60 015-Nov-2010 18-Nov-2010 18-Nov-2010 02-Dec-2010 02-Dec-2010 06-Dec-2010 16-Dec-2010 10-Jan-2011 Cumulative Sum of U_Result - Expected Value ( Value - Expected U_Result of Sum Cumulative DateReported
U_Result M ean of Cumulative Sum of U_Result - Expected Value (ppm) = -331.00
Printed: 03-Feb-2012 13:16:01 Data Imported: 02-Feb-2012 10:33:23 Page 1
Appendix A – QAQC Summary Plots Page: 1 Berlin Project Uranium Standards Feb 2012
Berlin Project Standards Feb 2012 (Standard ST1000020 ME-MS61U Analysis )
Standard: ST1000020 No of Analyses: 45 Element: U Res ult Minimum: 1,540.00 Units : ppm Max imum: 2,000.00 Detec tion Limit: 1 Mean: 1,802.44 Expected Value (EV): 1,910.20 Std Deviation: 83.91 E.V. Range: 1,719.18 to 2,101.22 % in Tolerance 88.89 % % Bias -5.64 % % RSD 4.66 %
Standard Control Plot (Standard ST1000020 ME-MS61U Analysis )
2200 2100 2000 1900 1800 1700
U_Result (ppm) U_Result 1600
1500 02-Feb-2011 09-Apr-2011 -2011 01-Jul 18-Aug-2011
DateReported
U_Result Expected Value = 1,910.20 EV Range (1,719.18 to 2,101.22) Mean of U_Result = 1,802.44
Cumulative Deviation from Assay Mean (Standard ST1000020 ME-MS61U Analysis )
100 0 -10 0 -20 0 -30 0
-40 0 02-Feb-2011 09-Apr-2011 -2011 01-Jul 18-Aug-2011 Cumulative Sum of U_Result - Mean (ppm) - Mean U_Result of Sum Cumulative
DateReported
U_Result Mean of Cumulative Sum of U_Result - Mean (ppm) = -148. 67 p
Cumulative Deviation from Expected Value (Standard ST1000020 ME-MS61U Analysis )
0 -10 00 -20 00 -30 00 -40 00
-50 00 02-Feb-2011 09-Apr-2011 -2011 01-Jul 18-Aug-2011 Cumulative Sum of U_Result - Expected Value ( Value - Expected U_Result of Sum Cumulative DateReported
U_Result Mean of Cumulative Sum of U_Result - Expected Value (ppm) = -2,627. 04
Printed: 03-Feb-2012 13:17:32 Data Imported: 02-Feb-2012 10:33:23 Page 1
Appendix A – QAQC Summary Plots Page: 2 Berlin Project Uranium Standards Feb 2012
Berlin Project Standards Feb 2012 (Standard ST1000045 ME-MS61U Analysis )
Standard: ST1000045 No of Analyses: 41 Element: U Res ult Minimum: 1,190.00 Units : ppm Max imum: 1,540.00 Detection Limit: 1 Mean: 1,259.76 Expected Value (EV): 1,353.00 Std Deviation: 69.23 E.V. Range: 1,217.70 to 1,488.30 % in Tolerance 82.93 % % Bias -6.89 % % RSD 5.50 %
Standard Control Plot (Standard ST1000045 ME-MS61U Analysis )
1600 1500 1400 1300 1200 U_Result (ppm) U_Result
1100 17-May-2011 -2011 01-Jul -2011 30-Jul 26-Sep-2011
DateReported
U_Result Expected Value = 1,353.00 EV Range (1,217.70 to 1,488.30) Mean of U_Result = 1,259.76
Cumulative Deviation from Assay Mean (Standard ST1000045 ME-MS61U Analysis )
200 0 -20 0 -40 0 -60 0
-80 0 17-May-2011 -2011 01-Jul -2011 30-Jul 26-Sep-2011 Cumulative Sum of U_Result - Mean (ppm) Mean - U_Result of Sum Cumulative
DateReported
U_Result Mean of Cumulative Sum of U_Result - Mean (ppm) = -334. 39 p
Cumulative Deviation from Expected Value (Standard ST1000045 ME-MS61U Analysis )
0 -10 00 -20 00 -30 00 -40 00
-50 00 17-May-2011 -2011 01-Jul -2011 30-Jul 26-Sep-2011 Cumulative Sum of U_Result - Expected Value ( Value - Expected U_Result of Sum Cumulative DateReported
U_Result M ean of Cumulative Sum of U_Result - Expected Value (ppm) = -2,292. 51
Printed: 03-Feb-2012 13:18:01 Data Imported: 02-Feb-2012 10:33:23 Page 1
Appendix A – QAQC Summary Plots Page: 3 Berlin Project Uranium Standards Feb 2012
Berlin Project Standards Feb 2012 (Standard AMIS-055 ME-MS61U Analysis )
Standard: AMIS-055 No of Analyses: 14 Element: U Result Minimum: 3,200.00 Units : ppm Max imum: 3,620.00 Detection Limit: 1 Mean: 3,360.71 Expected Value (EV): 3,206.00 Std Deviation: 112.15 E.V. Range: 2,885.40 to 3,526.60 % in Tolerance 92.86 % % Bias 4.83 % % RSD 3.34 %
Standard Control Plot (Standard AMIS-055 ME-MS61U Analysis )
3800 3600 3400 3200 3000 U_Result (ppm) U_Result
280020-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
DateReported
U_Result Expected Value = 3,206.00 EV Range (2,885.40 to 3,526.60) M ean of U_Result = 3,360.71
Cumulative Deviation from Assay Mean (Standard AMIS-055 ME-MS61U Analysis )
200 100 0 -10 0 -20 0 -30 0
-40 020-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 Cumulative Sum of U_Result - Mean (ppm) Mean - U_Result of Sum Cumulative
DateReported
U_Result Mean of Cumulative Sum of U_Result - Mean (ppm) = -45. 36 p Cumulative Deviation from Expected Value (Standard AMIS-055 ME-MS61U Analysis )
2500 2000 1500 1000 500 0
-50 -2011 020-Jul -2011 30-Jul 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 Cumulative Sum of U_Result - Expected Value ( Value - Expected U_Result of Sum Cumulative DateReported
U_Result Mean of Cumulative Sum of U_Result - Expected Value (ppm) = 1,115.00
Printed: 03-Feb-2012 13:19:25 Data Imported: 02-Feb-2012 10:33:23 Page 1
Appendix A – QAQC Summary Plots Page: 4 Berlin Project Vanadium Standards Feb 2012
Berlin Project Standards Feb 2012 (Standard ST1000020 V ME- MS61 Analysis)
Standard: ST1000020 No of Analyses: 9 Element: V Result Minimum: 3,100.00 Units : ppm Max imum: 3,490.00 Detection Limit: Mean: 3,274.44 Expected Value (EV): 3,414.00 Std Deviation: 141.27 E.V. Range: 3,072.60 to 3,755.40 % in Tolerance 100.00 % % Bias -4.09 % % RSD 4.31 %
Standard Control Plot (Standard ST1000020 V ME- MS61 Analysis)
3800
3600
3400
3200 V_Result (ppm) V_Result
300015-Nov-2010 18-Nov-2010 18-Nov-2010 02-Dec-2010 02-Dec-2010 06-Dec-2010 16-Dec-2010 10-Jan-2011
DateReported
V_Result Expected Value = 3,414.00 EV Range (3,072.60 to 3,755.40) M ean of V_Result = 3,274.44
Cumulative Deviation from Assay Mean (Standard ST1000020 V ME- MS61 Analysis)
100 0 -10 0 -20 0 -30 0
-40 015-Nov-2010 18-Nov-2010 18-Nov-2010 02-Dec-2010 02-Dec-2010 06-Dec-2010 16-Dec-2010 10-Jan-2011 Cumulative Sum of V_Result -(ppm) Mean V_Result of Sum Cumulative
DateReported
V_Result Mean of Cumulative Sum of V_Result - Mean (ppm) = -145.56 p Cumulative Deviation from Expected Value (Standard ST1000020 V ME- MS61 Analysis)
0 -20 0 -40 0 -60 0 -80 0 -10 00 -12 00
-14 0015-Nov-2010 18-Nov-2010 18-Nov-2010 02-Dec-2010 02-Dec-2010 06-Dec-2010 16-Dec-2010 10-Jan-2011 Cumulative Sum of V_Result - Expected Value ( Value - Expected V_Result of Sum Cumulative DateReported
V_Result Mean of Cumulative Sum of V_Result - Expected Value (ppm) = -843.33
Printed: 03-Feb-2012 11:52:35 Data Imported: 02-Feb-2012 12:54:00 Page 1
Appendix A – QAQC Summary Plots Page: 5 Berlin Project Vanadium Standards Feb 2012
Berlin Project Standards Feb 2012 (Standard ST1000020 V ME- MS61U Analysis )
Standard: ST1000020 No of Analyses: 44 Element: V Result Minimum: 3,040.00 Units: ppm Max imum: 3,420.00 Detection Limit: Mean: 3,203.86 Expected Value (EV): 3,414.00 Std Deviation: 95.83 E.V. Range: 3,072.60 to 3,755.40 % in Tolerance 86.36 % % Bias -6.16 % % RSD 2.99 %
Standard Control Plot (Standard ST1000020 V ME- MS61U Analysis )
3800
3600
3400
3200 V_Result (ppm)
3000 02-Feb-2011 09-Apr-2011 -2011 01-Jul 18-Aug-2011
DateReported
V_Result Expected Value = 3,414.00 EV Range (3,072.60 to 3,755.40) Mean of V_Result = 3,203.86
Cumulative Deviation from Assay Mean (Standard ST1000020 V ME- MS61U Analysis )
400 200 0 -20 0 -40 0
-60 0 02-Feb-2011 09-Apr-2011 01-Jul-2011 18-Aug-2011 Cumulative Sum of V_Result - Mean (ppm) - Mean V_Result of Sum Cumulative
DateReported
V_Result Mean of Cumulative Sum of V_Result - Mean (ppm) = 18.75 p Cumulative Deviation from Expected Value (Standard ST1000020 V ME- MS61U Analysis )
0 -20 00 -40 00 -60 00 -80 00
-10 000 02-Feb-2011 09-Apr-2011 01-Jul-2011 18-Aug-2011 Cumulative Sum of V_Result - Expected Value ( -Expected V_Result of Sum Cumulative DateReported
V_Result Mean of Cumulative Sum of V_Result - Expected Value (ppm) = -4,709. 32
Printed: 03-Feb-2012 11:50:02 Data Imported: 02-Feb-2012 12:54:00 Page 1
Appendix A – QAQC Summary Plots Page: 6 Berlin Project Vanadium Standards Feb 2012
U308Corp Berlin Project (Standard ST1000020 V ME- MS81Analysis )
Standard: ST1000020 No of Analyses: 33 Element: V Result Minimum: 2,950.00 Units: ppm Max imum: 3,950.00 Detec tion Limit: Mean: 3,393.94 Expected Value (EV): 3,414.00 Std Deviation: 256.01 E.V. Range: 3,072.60 to 3,755.40 % in Tolerance 75.76 % % Bias -0.59 % % RSD 7.54 %
Standard Control Plot (Standard ST1000020 V ME- MS81Analysis )
4000 3800 3600 3400 3200
V_Result (ppm) V_Result 3000
2800 11-Jan-2011 28-Feb-2011 19-Apr-2011
DateReported
V_Result Expected Value = 3,414.00 EV Range (3,072.60 to 3,755.40) Mean of V_Result = 3,393.94
Cumulative Deviation from Assay Mean (Standard ST1000020 V ME- MS81Analysis )
1000 0 -10 00 -20 00 -30 00
-40 00 11-Jan-2011 28-Feb-2011 19-Apr-2011 Cumulative Sum of V_Result - Mean (ppm) - Mean V_Result of Sum Cumulative
DateReported
V_Result Mean of Cumulative Sum of V_Result - Mean (ppm) = -1, 560.00 p Cumulative Deviation from Expected Value (Standard ST1000020 V ME- MS81Analysis )
0
-10 00
-20 00
-30 00
-40 00 11-Jan-2011 28-Feb-2011 19-Apr-2011
Cumulative Sum of V_Result - Expected Value ( Value - Expected V_Result of Sum Cumulative DateReported
V_Result Mean of Cumulative Sum of V_Result - Expected Value (ppm) = -1,901. 03
Printed: 02-Feb-2012 14:54:06 Data Imported: 02-Feb-2012 12:54:00 Page 1
Appendix A – QAQC Summary Plots Page: 7 Berlin Project Vanadium Standards Feb 2012
Berlin Project Standards Feb 2012 (Standard ST1000045 V ME- MS61U Analysis )
Standard: ST1000045 No of Analyses: 41 Element: V_Result Minimum: 3,020.00 Units: ppm Max imum: 3,580.00 Detection Limit: Mean: 3,348.54 Expected Value (EV): 3,584.00 Std Deviation: 106.49 E.V. Range: 3,225.60 to 3,942.40 % in Tolerance 90.24 % % Bias -6.57 % % RSD 3.18 %
Standard Control Plot (Standard ST1000045 V ME- MS61U Analysis )
4000 3800 3600 3400 3200 V_Result (ppm)
3000 17-May-2011 01-Jul-2011 30-Jul-2011 26-Sep-2011
DateReported
V_Result Expected Value = 3,584.00 EV Range (3,225.60 to 3,942.40) M ean of V_Result = 3,348.54
Cumulative Deviation from Assay Mean (Standard ST1000045 V ME- MS61U Analysis )
100 0 -10 0 -20 0 -30 0 -40 0 -50 0
-60 0 17-May-2011 01-Jul-2011 30-Jul-2011 26-Sep-2011 Cumulative Sum of V_Result - Mean (ppm) - Mean V_Result of Sum Cumulative
DateReported
V_Result Mean of Cumulative Sum of V_Result - Mean (ppm) = -187. 07 p
Cumulative Deviation from Expected Value (Standard ST1000045 V ME- MS61U Analysis )
0 -20 00 -40 00 -60 00 -80 00
-10 000 17-May-2011 -2011 01-Jul -2011 30-Jul 26-Sep-2011 Cumulative Sum of V_Result - Expected Value ( Value - Expected V_Result of Sum Cumulative DateReported
V_Result Mean of Cumulative Sum of V_Result - Expected Value (ppm) = -5,131.80
Printed: 03-Feb-2012 11:45:54 Data Imported: 02-Feb-2012 12:54:00 Page 1
Appendix A – QAQC Summary Plots Page: 8 Berlin Project Vanadium Standards Feb 2012
Berlin Project Standards Feb 2012 (Standard ST1000045 V ME- MS81 Analysis )
Standard: ST1000045 No of Analyses: 9 Element: V_Result Minimum: 3,320.00 Units : ppm Max imum: 4,030.00 Detection Limit: Mean: 3,740.00 Expected Value (EV): 3,584.00 Std Deviation: 233.71 E.V. Range: 3,225.60 to 3,942.40 % in Tolerance 77.78 % % Bias 4.35 % % RSD 6.25 %
Standard Control Plot (Standard ST1000045 V ME- MS81 Analysis )
4200 4000 3800 3600 3400 V_Result (ppm) V_Result
320023-Feb-2011 19-Apr-2011 19-Apr-2011 19-Apr-2011 21-Apr-2011 17-May-2011 26-May-2011 18-Jul-2011
DateReported
V_Result Expected Value = 3,584.00 EV Range (3,225.60 to 3,942.40) M ean of V_Result = 3,740.00
Cumulative Deviation from Assay Mean (Standard ST1000045 V ME- MS81 Analysis )
600 400 200 0 -20 0
-40 023-Feb-2011 19-Apr-2011 19-Apr-2011 19-Apr-2011 21-Apr-2011 17-May-2011 26-May-2011 18-Jul-2011 Cumulative Sum of V_Result - Mean (ppm) - Mean V_Result of Sum Cumulative
DateReported
V_Result Mean of Cumulative Sum of V_Result - Mean (ppm) = 28.89 p
Cumulative Deviation from Expected Value (Standard ST1000045 V ME- MS81 Analysis )
2000 1500 1000 500 0
-50 023-Feb-2011 19-Apr-2011 19-Apr-2011 19-Apr-2011 21-Apr-2011 17-May-2011 26-May-2011 -2011 18-Jul Cumulative Sum of V_Result - Expected Value ( Value - Expected V_Result of Sum Cumulative DateReported
V_Result M ean of Cumulative Sum of V_Result - Expected Value (ppm) = 808.89
Printed: 02-Mar-2012 15:21:23 Data Imported: 02-Feb-2012 12:54:00 Page 1
Appendix A – QAQC Summary Plots Page: 9 Berlin Project Molybdenum Standards Feb 2012
Berlin Project Standards Feb 2012 (Standard ST1000020 Mo ME-MS61U Analysis)
Standard: ST1000020 No of Analyses: 41 Element: Mo Res ult Minimum: 47.90 Units: ppm Max imum: 70.50 Detection Limit: Mean: 57.85 Expected Value (EV): 64.18 Std Deviation: 4.70 E.V. Range: 57.76 to 70.60 % in Tolerance 56.10 % % Bias -9.86 % % RSD 8.12 %
Standard Control Plot (Standard ST1000020 Mo ME-MS61U Analysis)
80
70
60 50 Mo_Result (ppm) Mo_Result
40 17-May-2011 01-Jul-2011 30-Jul-2011 26-Sep-2011
DateReported
Mo_Result Expected Value = 64.18 EV Range (57.76 to 70. 60) Mean of Mo_Result = 57.85 )
Cumulative Deviation from Assay Mean (Standard ST1000020 Mo ME-MS61U Analysis)
10 0 -10 -20 -30 -40
-50 17-May-2011 01-Jul-2011 30-Jul-2011 26-Sep-2011 Cumulative Sum of Mo_Result - Mean (ppm Mean - Mo_Result of Sum Cumulative
DateReported
Mo_Result Mean of Cumulative Sum of Mo_Result - Mean (ppm) = -16.48 ( Cumulative Deviation from Expected Value (Standard ST1000020 Mo ME-MS61U Analysis)
0
-10 0
-20 0
-30 0 17-May-2011 -2011 01-Jul -2011 30-Jul 26-Sep-2011
Cumulative Sum of Mo_Result - Expected Value Value - Expected Mo_Result of Sum Cumulative DateReported
Mo_Result Mean of Cumulative Sum of Mo_Result - Expected Value (ppm ) = -149.44
Printed: 03-Feb-2012 11:32:58 Data Imported: 02-Feb-2012 15:00:02 Page 1
Appendix A – QAQC Summary Plots Page: 10 Berlin Project Molybdenum Standards Feb 2012
Berlin Project Standards Feb 2012 (Standard ST1000020 Mo ME-MS81 Analysis)
Standard: ST1000020 No of Analyses: 44 Element: Mo_Result Minimum: 52.90 Units : ppm Max imum: 76.40 Detection Limit: Mean: 65.59 Expected Value (EV): 64.18 Std Deviation: 4.21 E.V. Range: 57.76 to 70.60 % in Tolerance 84.09 % % Bias 2.20 % % RSD 6.42 %
Standard Control Plot (Standard ST1000020 Mo ME-MS81 Analysis)
80
70
60 Mo_Result (ppm) Mo_Result
50 02-Feb-2011 09-Apr-2011 -2011 01-Jul 18-Aug-2011
DateReported
Mo_Result Expected Value = 64.18 EV Range (57.76 to 70. 60) Mean of Mo_Result = 65.59 ) Cumulative Deviation from Assay Mean (Standard ST1000020 Mo ME-MS81 Analysis)
10 0 -10 -20 -30
-40 02-Feb-2011 09-Apr-2011 -2011 01-Jul 18-Aug-2011 Cumulative Sum of Mo_Result - Mean (ppm Mean - Mo_Result of Sum Cumulative
DateReported
Mo_Result Mean of Cumulative Sum of Mo_Result - Mean (ppm) = -15.67 ( Cumulative Deviation from Expected Value (Standard ST1000020 Mo ME-MS81 Analysis)
80 60 40 20 0
-20 02-Feb-2011 09-Apr-2011 01-Jul-2011 18-Aug-2011
Cumulative Sum of Mo_Result - Expected Value Value Expected - Mo_Result of Sum Cumulative DateReported
Mo_Result Mean of Cumulative Sum of Mo_Result - Expected Value (ppm) = 16.08
Printed: 03-Feb-2012 11:32:30 Data Imported: 02-Feb-2012 15:00:02 Page 1
Appendix A – QAQC Summary Plots Page: 11 Berlin Project Molybdenum Standards Feb 2012
Berlin Project Standards Feb 2012 (Standard ST1000020 Mo ME-MS61 Analysis )
Standard: ST1000020 No of Analyses: 9 Element: Mo_Result Minimum: 61.20 Units : ppm Max imum: 70.80 Detection Limit: Mean: 66.71 Expected Value (EV): 64.18 Std Deviation: 3.36 E.V. Range: 57.76 to 70.60 % in Tolerance 88.89 % % Bias 3.94 % % RSD 5.03 %
Standard Control Plot (Standard ST1000020 Mo ME-MS61 Analysis )
75
70
65
60 Mo_Result (ppm) Mo_Result
5515-Nov-2010 18-Nov-2010 18-Nov-2010 02-Dec-2010 02-Dec-2010 06-Dec-2010 16-Dec-2010 10-Jan-2011
DateReported
Mo_Result Expected Value = 64.18 EV Range (57.76 to 70. 60) Mean of Mo_Result = 66.71 ) Cumulative Deviation from Assay Mean (Standard ST1000020 Mo ME-MS61 Analysis )
4 2 0 -2 -4 -6
-815-Nov-2010 18-Nov-2010 18-Nov-2010 02-Dec-2010 02-Dec-2010 06-Dec-2010 16-Dec-2010 10-Jan-2011 Cumulative Sum of Mo_Result - Mean (ppm Mean - Mo_Result of Sum Cumulative
DateReported
Mo_Result Mean of Cumulative Sum of Mo_Result - Mean (ppm) = -1.41 ( Cumulative Deviation from Expected Value (Standard ST1000020 Mo ME-MS61 Analysis )
25
20
15
10
515-Nov-2010 18-Nov-2010 18-Nov-2010 02-Dec-2010 02-Dec-2010 06-Dec-2010 16-Dec-2010 10-Jan-2011
Cumulative Sum of Mo_Result - Expected Value Value Expected - Mo_Result of Sum Cumulative DateReported
Mo_Result Mean of Cumulative Sum of Mo_Result - Expected Value (ppm) = 11.24
Printed: 03-Feb-2012 11:31:23 Data Imported: 02-Feb-2012 15:00:02 Page 1
Appendix A – QAQC Summary Plots Page: 12 Berlin Project Phosphorus Standards Feb 2012
Berlin Project Standards Feb 2012 (Standard ST1000020 P ME-MS81 Analysis)
Standard: ST1000020 No of Analyses: 33 Element: P Res ult Minimum: 1.80 Units: % Max imum: 1.94 Detection Limit: Mean: 1.87 Expected Value (EV): 1.76 Std Deviation: 0.03 E.V. Range: 1.58 to 1.94 % in Tolerance 96.97 % % Bias 6.23 % % RSD 1.76 %
Standard Control Plot (Standard ST1000020 P ME-MS81 Analysis)
1.9 1.8 1.7
P_Result (%) 1.6
1.5 11-Jan-2011 28-Feb-2011 19-Apr-2011
DateReported
P_Result Expected Value = 1.76 EV Range (1.58 to 1. 94) M ean of P_Result = 1.87
Cumulative Deviation from Assay Mean (Standard ST1000020 P ME-MS81 Analysis)
0.05 0.00 -0.05 -0.10 -0.15 -0.20
-0.25 11-Jan-2011 28-Feb-2011 19-Apr-2011 Cumulative Sum of P_Result - Mean (%) Mean - P_Result of Sum Cumulative
DateReported
P_Result Mean of Cumulative Sum of P_Result - Mean (%) = -0.12
Cumulative Deviation from Expected Value (Standard ST1000020 P ME-MS81 Analysis)
4
3
2
1
0 11-Jan-2011 28-Feb-2011 19-Apr-2011 Cumulative Sum of P_Result - Expected Value ( Value - Expected P_Result of Sum Cumulative DateReported
P_Result Mean of Cumulative Sum of P_Result - Expected Value (%) = 1.74
Printed: 03-Feb-2012 11:29:09 Data Edited: 02-Feb-2012 17:00:20 Page 1
Appendix A – QAQC Summary Plots Page: 13 Berlin Project Phosphorus Standards Feb 2012
Berlin Project Standards Feb 2012 (Standard ST1000020 P ME-XRF12 Analysis )
Standard: ST1000020 No of Analyses: 37 Element: P Res ult Minimum: 1.80 Units : % Max imum: 1.94 Detec tion Limit: Mean: 1.87 Expected Value (EV): 1.76 Std Deviation: 0.03 E.V. Range: 1.58 to 1.94 % in Tolerance 97.30 % % Bias 6.36 % % RSD 1.75 %
Standard Control Plot (Standard ST1000020 P ME-XRF12 Analysis )
1.9 1.8 1.7
P_Result (%) P_Result 1.6
1.5 27-Jan-2011 12-Mar-2011 08-Jun-2011
DateReported
P_Result Expected Value = 1.76 EV Range (1.58 to 1. 94) Mean of P_Result = 1.87
Cumulative Deviation from Assay Mean (Standard ST1000020 P ME-XRF12 Analysis )
0.05 0.00 -0.05 -0.10 -0.15
-0.20 27-Jan-2011 12-Mar-2011 08-Jun-2011 Cumulative Sum of P_Result - Mean (%) - Mean P_Result of Sum Cumulative
DateReported
P_Result Mean of Cumulative Sum of P_Result - Mean (%) = -0.07 ( Cumulative Deviation from Expected Value (Standard ST1000020 P ME-XRF12 Analysis )
5 4 3 2 1
0 27-Jan-2011 12-Mar-2011 08-Jun-2011 Cumulative Sum of P_Result - Expected Value Value Expected - P_Result of Sum Cumulative DateReported
P_Result Mean of Cumulative Sum of P_Result - Expected Value (%) = 2.06
Printed: 03-Feb-2012 11:28:37 Data Edited: 02-Feb-2012 17:00:20 Page 1
Appendix A – QAQC Summary Plots Page: 14 Berlin Project Phosphorus Standards Feb 2012
Berlin Project Standards Feb 2012 (Standard ST1000045 P ME-MS81 Analysis )
Standard: ST1000045 No of Analyses: 9 Element: P_Res ult Minimum: 2.45 Units: % Max imum: 2.54 Detection Limit: Mean: 2.49 Expected Value (EV): 2.44 Std Deviation: 0.02 E.V. Range: 2.20 to 2.68 % in Tolerance 100.00 % % Bias 2.10 % % RSD 0.94 %
Standard Control Plot (Standard ST1000045 P ME-MS81 Analysis )
2.6 2.5 2.4 2.3
P_Result (%) P_Result 2.2
2.123-Feb-2011 19-Apr-2011 19-Apr-2011 19-Apr-2011 21-Apr-2011 17-May-2011 26-May-2011 -2011 18-Jul
DateReported
P_Result Expected Value = 2.44 EV Range (2.20 to 2. 68) M ean of P_Result = 2.49
Cumulative Deviation from Assay Mean (Standard ST1000045 P ME-MS81 Analysis )
0.01 0.00 -0.01 -0.02 -0.03 -0.04
-0.0523-Feb-2011 19-Apr-2011 19-Apr-2011 19-Apr-2011 21-Apr-2011 17-May-2011 26-May-2011 18-Jul-2011 Cumulative Sum of P_Result - Mean (%) Mean - P_Result of Sum Cumulative
DateReported
P_Result Mean of Cumulative Sum of P_Result - Mean (%) = -0. 01 ( Cumulative Deviation from Expected Value (Standard ST1000045 P ME-MS81 Analysis )
0.5 0.4 0.3 0.2 0.1
0.023-Feb-2011 19-Apr-2011 19-Apr-2011 19-Apr-2011 21-Apr-2011 17-May-2011 26-May-2011 18-Jul-2011 Cumulative Sum of P_Result - Expected Value Value Expected - P_Result of Sum Cumulative DateReported
P_Result Mean of Cumulative Sum of P_Result - Expected Value (%) = 0.24
Printed: 03-Feb-2012 11:27:27 Data Edited: 02-Feb-2012 17:00:20 Page 1
Appendix A – QAQC Summary Plots Page: 15 Berlin Project Phosphorus Standards Feb 2012
Berlin Project Standards Feb 2012 (Standard ST1000045 P ME-XRF12 Analysis )
Standard: ST1000045 No of Analyses: 20 Element: P_Res ult Minimum: 2.45 Units: % Max imum: 2.54 Detection Limit: Mean: 2.49 Expected Value (EV): 2.44 Std Deviation: 0.02 E.V. Range: 2.20 to 2.68 % in Tolerance 100.00 % % Bias 1.94 % % RSD 0.83 %
Standard Control Plot (Standard ST1000045 P ME-XRF12 Analysis )
2.6 2.5 2.4 2.3
P_Result (%) 2.2
2.123-Feb-2011 19-Apr-2011 19-Apr-2011 19-Apr-2011 21-Apr-2011 07-May-2011 16-May-2011 16-May-2011 17-May-2011 17-May-2011 26-May-2011 08-Jun-2011 -2011 07-Jul -2011 08-Jul -2011 18-Jul -2011 18-Jul -2011 18-Jul -2011 20-Jul 25-Aug-2011
DateReported
P_Result Expected Value = 2.44 EV Range (2.20 to 2. 68) M ean of P_Result = 2.49
Cumulative Deviation from Assay Mean (Standard ST1000045 P ME-XRF12 Analysis )
0.02 0.00 -0.02 -0.04
-0.0623-Feb-2011 19-Apr-2011 19-Apr-2011 19-Apr-2011 21-Apr-2011 07-May-2011 16-May-2011 16-May-2011 17-May-2011 17-May-2011 26-May-2011 08-Jun-2011 07-Jul-2011 08-Jul-2011 18-Jul-2011 18-Jul-2011 18-Jul-2011 20-Jul-2011 25-Aug-2011 Cumulative Sum of P_Result - Mean (%) - Mean P_Result of Sum Cumulative
DateReported
P_Result Mean of Cumulative Sum of P_Result - Mean (%) = -0.01 ( Cumulative Deviation from Expected Value (Standard ST1000045 P ME-XRF12 Analysis )
1.0 0.8 0.6 0.4 0.2
0.023-Feb-2011 19-Apr-2011 19-Apr-2011 19-Apr-2011 21-Apr-2011 07-May-2011 16-May-2011 16-May-2011 17-May-2011 17-May-2011 26-May-2011 08-Jun-2011 -2011 07-Jul -2011 08-Jul -2011 18-Jul -2011 18-Jul -2011 18-Jul -2011 20-Jul 25-Aug-2011 Cumulative Sum of P_Result - Expected Value Value Expected - P_Result of Sum Cumulative DateReported
P_Result Mean of Cumulative Sum of P_Result - Expected Value (%) = 0.48
Printed: 03-Feb-2012 11:27:57 Data Edited: 02-Feb-2012 17:00:20 Page 1
Appendix A – QAQC Summary Plots Page: 16 Berlin Project Phosphorus Standards Feb 2012
Berlin Project Standards Feb 2012 (Standard AMIS-055 P ME-MS81 Analysis)
Standard: AMIS-055 No of Analyses: 8 Element: P_Res ult Minimum: 9.21 Units: Max imum: 9.64 Detection Limit: Mean: 9.36 Expected Value (EV): 9.36 Std Deviation: 0.12 E.V. Range: 8.42 to 10.30 % in Tolerance 100.00 % % Bias 0.01 % % RSD 1.32 %
Standard Control Plot (Standard AMIS-055 P ME-MS81 Analysis)
10.5 10.0 9.5 9.0
P_Result (%) P_Result 8.5
8.0 -2011 20-Jul 19-Aug-2011 26-Aug-2011 05-Sep-2011 19-Sep-2011 01-Oct-2011 01-Oct-2011
DateReported
P_Result Expected Value = 9.36 EV Range (8.42 to 10. 30) Mean of P_Result = 9.36
Cumulative Deviation from Assay Mean (Standard AMIS-055 P ME-MS81 Analysis)
0.10
0.00
-0.10
-0.20 -2011 20-Jul 19-Aug-2011 26-Aug-2011 05-Sep-2011 19-Sep-2011 01-Oct-2011 01-Oct-2011 Cumulative Sum of P_Result - Mean (%) - Mean P_Result of Sum Cumulative
DateReported
P_Result M ean of C um ul ative S um of P_R es ul t - M ean (% ) = -0. 04 ( Cumulative Deviation from Expected Value (Standard AMIS-055 P ME-MS81 Analysis)
0.10
0.00
-0.10
-0.2020-Jul-2011 19-Aug-2011 26-Aug-2011 05-Sep-2011 19-Sep-2011 01-Oct-2011 01-Oct-2011 Cumulative Sum of P_Result - Expected Value Value Expected - P_Result of Sum Cumulative DateReported
P_Result Mean of Cumulative Sum of P_Result - Expected Value (% ) = -0.04
Printed: 03-Feb-2012 11:23:05 Data Edited: 02-Feb-2012 17:00:20 Page 1
Appendix A – QAQC Summary Plots Page: 17 Berlin Project Phosphorus Standards Feb 2012
Berlin Project Standards Feb 2012 (Standard AMIS-055 P ME-XRF12 Analysis)
Standard: AMIS-055 No of Analyses: 9 Element: P_Res ult Minimum: 9.21 Units: Max imum: 9.64 Detection Limit: Mean: 9.36 Expected Value (EV): 9.36 Std Deviation: 0.12 E.V. Range: 8.42 to 10.30 % in Tolerance 100.00 % % Bias -0.02 % % RSD 1.25 %
Standard Control Plot (Standard AMIS-055 P ME-XRF12 Analysis)
10.5 10.0 9.5 9.0
P_Result (%) P_Result 8.5
8.018-Aug-2011 19-Aug-2011 26-Aug-2011 05-Sep-2011 19-Sep-2011 19-Sep-2011 01-Oct-2011 01-Oct-2011
DateReported
P_Result Expected Value = 9.36 EV Range (8.42 to 10. 30) Mean of P_Result = 9.36
Cumulative Deviation from Assay Mean (Standard AMIS-055 P ME-XRF12 Analysis)
0.10
0.00
-0.10
-0.2018-Aug-2011 19-Aug-2011 26-Aug-2011 05-Sep-2011 19-Sep-2011 19-Sep-2011 01-Oct-2011 01-Oct-2011 Cumulative Sum of P_Result - Mean (%) - Mean P_Result of Sum Cumulative
DateReported
P_Result M ean of C um ul ative S um of P_R es ul t - M ean (% ) = -0. 04 ( Cumulative Deviation from Expected Value (Standard AMIS-055 P ME-XRF12 Analysis)
0.10
0.00
-0.10
-0.2018-Aug-2011 19-Aug-2011 26-Aug-2011 05-Sep-2011 19-Sep-2011 19-Sep-2011 01-Oct-2011 01-Oct-2011 Cumulative Sum of P_Result - Expected Value Value Expected - P_Result of Sum Cumulative DateReported
P_Result Mean of Cumulative Sum of P_Result - Expected Value (% ) = -0.05
Printed: 03-Feb-2012 11:18:46 Data Edited: 02-Feb-2012 17:00:20 Page 1
Appendix A – QAQC Summary Plots Page: 18 Berlin Project Yttrium Standards Feb 2012
Berlin Project Standards Feb 2012 (Standard AMIS-055 ME-MS61U Analysis)
AMIS-055 Y ME-MS81 Standard: Analysis No of Analyses: 14 Element: Y Result Minimum: 51.20 Units: ppm Max imum: 61.40 Detection Limit: Mean: 56.29 Expected Value (EV): 53.60 Std Deviation: 3.01 E.V. Range: 48.24 to 58.96 % in Tolerance 71.43 % % Bias 5.01 % % RSD 5.35 %
Standard Control Plot (Standard AMIS-055 ME-MS61U Analysis)
62 60 58 56 54 52 50 Y_Result (ppm) Y_Result
4820-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
DateReported
Y_Result Expected Value = 53.60 EV Range (48.24 to 58.96) Mean of Y_Result = 56. 29
Cumulative Deviation from Assay Mean (Standard AMIS-055 ME-MS61U Analysis)
10 5 0 -5 -10
-1520-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 Cumulative Sum of Y_Result - Mean (ppm) - Mean Y_Result of Sum Cumulative
DateReported
Y_Result Mean of Cumulative Sum of Y_Result - Mean (ppm) = -0.97
Cumulative Deviation from Expected Value (Standard AMIS-055 ME-MS61U Analysis)
40 30 20 10 0
-1020-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
Cumulative Sum of Y_Result - Expected Value ( Value - Expected Y_Result of Sum Cumulative DateReported
Y_Result Mean of Cumulative Sum of Y_Result - Expected Value (ppm) = 19.17
Printed: 02-Mar-2012 15:25:51 Data Imported: 03-Feb-2012 09:42:32 Page 1
Appendix A – QAQC Summary Plots Page: 19 Berlin Project Yttrium Standards Feb 2012
Berlin Project Standards Feb 2012 (Standard AMIS-055 ME-MS81 Analysis )
Standard: AMIS-055 No of Analyses: 8 Element: Y Result Minimum: 51.50 Units: ppm Max imum: 58.90 Detection Limit: - Mean: 55.41 Expected Value (EV): 53.60 Std Deviation: 2.52 E.V. Range: 48.24 to 58.96 % in Tolerance 100.00 % % Bias 3.38 % % RSD 4.54 %
Standard Control Plot (Standard AMIS-055 ME-MS81 Analysis )
60 58 56 54 52 50 Y_Result (ppm) Y_Result
48 -2011 20-Jul 19-Aug-2011 26-Aug-2011 05-Sep-2011 19-Sep-2011 01-Oct-2011 01-Oct-2011
DateReported
Y_Result Expected Value = 53.60 EV Range (48.24 to 58. 96) M ean of Y_Result = 55.41
Cumulative Deviation from Assay Mean (Standard AMIS-055 ME-MS81 Analysis )
6 5 4 3 2 1 0
-1 -2011 20-Jul 19-Aug-2011 26-Aug-2011 05-Sep-2011 19-Sep-2011 01-Oct-2011 01-Oct-2011 Cumulative Sum of Y_Result - Mean (ppm) - Mean Y_Result of Sum Cumulative
DateReported
Y_Result Mean of Cumulative Sum of Y_Result - Mean (ppm) = 2.51 p Cumulative Deviation from Expected Value (Standard AMIS-055 ME-MS81 Analysis )
16 14 12 10 8 6
420-Jul-2011 19-Aug-2011 26-Aug-2011 05-Sep-2011 19-Sep-2011 01-Oct-2011 01-Oct-2011 Cumulative Sum of Y_Result - Expected Value ( Value Expected - Y_Result of Sum Cumulative DateReported
Y_Result Mean of Cumulative Sum of Y_Result - Expected Value (ppm) = 10.66
Printed: 03-Feb-2012 11:14:57 Data Imported: 03-Feb-2012 09:42:32 Page 1
Appendix A – QAQC Summary Plots Page: 20 Berlin Project U3O8 Corp. Blanks
Berlin Project Standards Feb 2012 (Blank ME-MS61U Analysis)
Standard: Blank No of Analyses: 63 Element: U Result Minimum: 0.90 Units: ppm Max imum: 18.90 Detection Limit: 1 Mean: 2.85 Expected Value (EV): 2.97 Std Deviation: 2.89 E.V. Range: 2.67 to 3.27 % in Tolerance 12.70 % % Bias -3.96 % % RSD 101.36 %
Standard Control Plot (Blank ME-MS61U Analysis)
20
15
10
5 U_Result (ppm) U_Result
0 10-Apr-2011 25-May-2011 -2011 02-Jul 04-Aug-2011 20-Aug-2011 01-Oct-2011
DateReported
U_Result Expected Value = 2.97 EV Range (2.67 to 3. 27) M ean of U_Result = 2.85
Cumulative Deviation from Assay Mean (Blank ME-MS61U Analysis)
0 -10 -20 -30 -40 -50
-60 10-Apr-2011 25-May-2011 -2011 02-Jul 04-Aug-2011 20-Aug-2011 01-Oct-2011 Cumulative Sum of U_Result - Mean (ppm) Mean - U_Result of Sum Cumulative
DateReported
U_Result M ean of Cumulative Sum of U_Result - Mean (ppm) = -33. 02 p
Cumulative Deviation from Expected Value (Blank ME-MS61U Analysis)
0 -10 -20 -30 -40 -50
-60 10-Apr-2011 25-May-2011 -2011 02-Jul 04-Aug-2011 20-Aug-2011 01-Oct-2011 Cumulative Sum of U_Result - Expected Value ( Value - Expected U_Result of Sum Cumulative DateReported
U_Result Mean of Cumulative Sum of U_Result - Expected Value (ppm) = -36.78
Printed: 03-Feb-2012 13:14:55 Data Imported: 02-Feb-2012 10:33:23 Page 1
Appendix A – QAQC Summary Plots Page: 21 Berlin Project U3O8 Corp. Blanks
Berlin Project Standards Feb 2012 (Blank ME-MS81 Analysis )
Standard: Blank No of Analyses: 24 Element: U Result Minimum: 1.31 Units: ppm Max imum: 11.65 Detection Limit: 1 Mean: 3.52 Expected Value (EV): 2.97 Std Deviation: 2.64 E.V. Range: 2.67 to 3.27 % in Tolerance 0.00 % % Bias 18.43 % % RSD 75.03 %
Standard Control Plot (Blank ME-MS81 Analysis )
12 10 8 6 4 2 U_Result (ppm) U_Result
002-Dec-2010 10-Jan-2011 11-Jan-2011 02-Feb-2011 02-Feb-2011 04-Feb-2011 28-Feb-2011 04-Mar-2011 12-Mar-2011 09-Apr-2011 10-Apr-2011 13-Apr-2011 19-Apr-2011 21-Apr-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
DateReported
U_Result Expected Value = 2.97 EV Range (2.67 to 3. 27) M ean of U_Result = 3.52
Cumulative Deviation from Assay Mean (Blank ME-MS81 Analysis )
0
-10
-20
-3002-Dec-2010 10-Jan-2011 11-Jan-2011 02-Feb-2011 02-Feb-2011 04-Feb-2011 28-Feb-2011 04-Mar-2011 12-Mar-2011 09-Apr-2011 10-Apr-2011 13-Apr-2011 19-Apr-2011 21-Apr-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 Cumulative Sum of U_Result - Mean (ppm) Mean - U_Result of Sum Cumulative
DateReported
U_Result M ean of Cumulative Sum of U_Result - Mean (ppm) = -12. 78 p Cumulative Deviation from Expected Value (Blank ME-MS81 Analysis )
20
10
0
-10
-2002-Dec-2010 10-Jan-2011 11-Jan-2011 02-Feb-2011 02-Feb-2011 04-Feb-2011 28-Feb-2011 04-Mar-2011 12-Mar-2011 09-Apr-2011 10-Apr-2011 13-Apr-2011 19-Apr-2011 21-Apr-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 Cumulative Sum of U_Result - Expected Value ( Value - Expected U_Result of Sum Cumulative DateReported
U_Result Mean of Cumulative Sum of U_Result - Expected Value (ppm) = -5.94
Printed: 03-Feb-2012 13:16:43 Data Imported: 02-Feb-2012 10:33:23 Page 1
Appendix A – QAQC Summary Plots Page: 22 Berlin Project U3O8 Corp. Duplicates
Berlin Project Duplicate Analysis Feb 12 (Trench Lab Pulp U ME-MS61 Analysis)
Orig_Assay Duplicate_A _U ssay_U Units Result No. Pairs: 4 4 Pearson CC: 1.00 Minimum: 7.70 8.00 ppm Spearman CC: 1.00 Max imum: 710.00 710.00 ppm Mean HA RD: 0.77 Mean: 243.68 243.00 ppm Median HA RD: 0.59 Median 128.50 127.00 ppm Std. Deviation: 273.88 274.13 ppm Mean HRD: -0.18 Coefficient of Variation: 1.12 1.13 Median HRD 0.27
Mean vs. HARD Plot Rank HARD Plot (Trench Lab Pulp U ME-MS61 Analysis) (Trench Lab Pulp U ME-MS61 Analysis)
10 100
5 50 HARD (%) HARD 0 (%) HARD 11010010000 0 102030405060708090100 Mean of Data Pair (ppm) Rank (%)
Mean HARD: 0.77 Median HARD: 0. 59 Precision: 10% Precision: 10%
HRD Histogram Mean vs. HRD Plot (Trench Lab Pulp U ME-MS61 Analysis) (Trench Lab Pulp U ME-MS61 Analysis)
100 10 5 50 0 -5 HRD (%) HRD 0 -10 Frequency (%) Frequency -1.0 0.0 1.0 1101001000 HRD (/100) Mean of Data Pair (ppm)
Mean HRD: -0.18 Median HRD: 0.27 Mean HRD: -0.18 Median HRD: 0.27 Precision: +/-10% Precision: +/-10%
T & H Precision Plot (Assay Pairs) T & H Precision Plot (Grouped Pairs) (Trench Lab Pulp U ME-MS61 Analysis) (Trench Lab Pulp U ME-MS61 Analysis)
1000 100 10 1 There is not enough data to generate this plot. 0.1 1101001000 Mean of Dat a Pair (ppm) Absolute Differen (ppm) ce
10% 20% 30% 50%
Correlation Plot QQ Plot (Trench Lab Pulp U ME-MS61 Analysis) (Trench Lab Pulp U ME-MS61 Analysis)
800 800 600 600 400 400 200 200 0 0 0 100 200 300 400 500 600 700 800 0 100 200 300 400 500 600 700 800 Orig_Assay_U (ppm) Orig_A ssay_U (ppm) Duplicate_Assay_U (ppm) Duplicate_Assay_U (ppm) Duplicate_Assay_U
P.CC= 1.00 S.CC= 1.00 Ref. Line y = 1.00x -0.90 Ref. Line y = 1.00x -0.90
Printed: 05-Feb-2012 21:22:03 Data Imported: 04-Feb-2012 21:45:53 Page 1
Appendix A – QAQC Summary Plots Page: 23 Berlin Project U3O8 Corp. Duplicates
Berlin Project Duplicate Analysis Feb 12 (Lab Pulp U ME-MS61 Analysis)
Orig_Assay Duplicate_A _U ssay_U Units Result No. Pairs: 4 4 Pearson CC: 1.00 Minimum: 7.70 8.00 ppm Spearman CC: 1.00 Max imum: 710.00 710.00 ppm Mean HA RD: 0.77 Mean: 243.68 243.00 ppm Median HARD: 0.59 Median 128.50 127.00 ppm Std. Deviation: 273.88 274.13 ppm Mean HRD: -0.18 Coefficient of Variation: 1.12 1.13 Median HRD 0.27
Mean vs. HARD Plot Rank HARD Plot (Lab Pulp U ME-MS61 Analysis) (Lab Pulp U ME-MS61 Analysis)
10 100
5 50 HARD (%) HARD 0 (%) HARD 1 10 100 1000 0 0 102030405060708090100 Mean of Data Pair (ppm) Rank (%)
M ean HARD: 0.77 M edian HARD: 0. 59 Precision: 10% Precision: 10%
HRD Histogram Mean vs. HRD Plot (Lab Pulp U ME-MS61 Analysis) (Lab Pulp U ME-MS61 Analysis)
100 10 5 50 0 -5 HRD (%) HRD 0 -10 Frequency (%) Frequency -1.0 0.0 1.0 1 10 100 1000 HRD (/100) Mean of Data Pair (ppm)
Mean HRD: -0.18 M edian HRD: 0.27 Mean HRD: -0.18 M edian HRD: 0.27 Precision: + /-10% Precision: + /-10%
T & H Precision Plot (Assay Pairs) T & H Precision Plot (Grouped Pairs) (Lab Pulp U ME-MS61 Analysis) (Lab Pulp U ME-MS61 Analysis)
1000 100 10 1 There is not enough data to generate this plot. 0.1 1 10 100 1000 Mean of Data Pair (ppm) Absolu te Difference (ppm)
10% 20% 30% 50%
Correlation Plot QQ Plot (Lab Pulp U ME-MS61 Analysis) (Lab Pulp U ME-MS61 Analysis)
800 800 600 600 400 400 200 200 0 0 0 100 200 300 400 500 600 700 800 0 100 200 300 400 500 600 700 800 Orig_A ssay_U (ppm) Orig_A ssay_U (ppm) Duplicate_Assay_U (ppm) Duplicate_Assay_U (ppm)
P.CC= 1.00 S.CC= 1.00 Ref. Line y = 1.00x -0.90 Ref. Line y = 1.00x -0.90
Printed: 05-Feb-2012 21:20:02 Data Imported: 04-Feb-2012 21:45:53 Page 1
Appendix A – QAQC Summary Plots Page: 24 Berlin Project U3O8 Corp. Duplicates
Berlin Project Duplicate Analysis Feb 12 (U3O8Corp 10 Mesh U ME-MS61U Analysis)
Orig_Assay Duplicate_A _U ssay_U Units Result No. Pairs: 10 10 Pearson CC: 1.00 Minimum: 5.30 3.80 ppm Spearman CC: 0.93 Max imum: 780.00 760.00 ppm Mean HA RD: 7.50 Mean: 132.51 131.72 ppm Median HA RD: 1.37 Median 13.95 17.95 ppm Std. Deviation: 239.99 234.07 ppm Mean HRD: -3.03 Coefficient of Variation: 1.81 1.78 Median HRD 0.34
Mean vs. HARD Plot Rank HARD Plot (U3O8Corp 10 Mesh U ME-MS61U Analysis) (U3O8Corp 10 Mesh U ME-MS61U Analysis)
60 100
40 80.00% of data are within 50 Precision limits 20 HARD (%) HARD 0 (%) HARD 11010010000 0 102030405060708090100 Mean of Data Pair (ppm) Rank (%)
Mean HARD: 7.50 Median HARD: 1.37 Precision: 10% Precision: 10%
HRD Histogram Mean vs. HRD Plot (U3O8Corp 10 Mesh U ME-MS61U Analysis) (U3O8Corp 10 Mesh U ME-MS61U Analysis)
80 20 60 0 40 -20 20 -40 HRD (%) HRD 0 -60 Frequency (%) Frequency -1.0 0.0 1.0 1101001000 HRD (/100) Mean of Data Pair (ppm)
Mean HRD: -3.03 Median HRD: 0.34 Mean HRD: -3.03 Median HRD: 0.34 Precision: +/-10% Precision: +/-10%
T & H Precision Plot (Assay Pairs) T & H Precision Plot (Grouped Pairs) (U3O8Corp 10 Mesh U ME-MS61U Analysis) (U3O8Corp 10 Mesh U ME-MS61U Analysis)
1000 100 10 1 There is not enough data to generate this plot. 0.1 1101001000 Mean of Dat a Pair (ppm) Absolu te Difference (ppm)
10% 20% 30% 50%
Correlation Plot QQ Plot (U3O8Corp 10 Mesh U ME-MS61U Analysis) (U3O8Corp 10 Mesh U ME-MS61U Analysis)
800 800 600 600 400 400 200 200 0 0 100 200 300 400 500 600 700 800 0 0 100 200 300 400 500 600 700 800 Orig_Assay_U (ppm)
Duplicate_Assay_U (ppm) Orig_A ssay_U (ppm) Duplicate_Assay_U (ppm)
P.CC= 1.00 S.CC= 0.93 Ref. Line y = 0.98x + 2. 51 Ref. Line y = 0.98x + 2. 50
Printed: 05-Feb-2012 21:07:47 Data Imported: 04-Feb-2012 21:45:53 Page 1
Appendix A – QAQC Summary Plots Page: 25 Berlin Project U3O8 Corp. Duplicates
Berlin Project Duplicate Analysis Feb 12 (U3O8Corp Sample DH U ME-MS61U Analysis)
Orig_Assay Duplicate_A _U ssay_U Units Result No. Pairs: 15 15 Pearson CC: 1.00 Minimum: 10.00 5.90 ppm Spearman CC: 0.99 Max imum: 1,010.00 970.00 ppm Mean HA RD: 4.35 Mean: 247.67 248.28 ppm Median HA RD: 2.40 Median 27.10 23.30 ppm Std. Deviation: 339.81 343.76 ppm Mean HRD: 1.69 Coefficient of Variation: 1.37 1.38 Median HRD 0.81
Mean vs. HARD Plot Rank HARD Plot (U3O8Corp Sample DH U ME-MS61U Analysis) (U3O8Corp Sample DH U ME-MS61U Analysis)
30 100
20 93.33% of data are within 50 Precision limits 10 HARD (%) HARD 0 (%) HARD 11010010000 0 102030405060708090100 Mean of Data Pair (ppm) Rank (%)
Mean HARD: 4.35 Median HARD: 2.40 Precision: 10% Precision: 10%
HRD Histogram Mean vs. HRD Plot (U3O8Corp Sample DH U ME-MS61U Analysis) (U3O8Corp Sample DH U ME-MS61U Analysis)
60 30
40 20 10 20 0 HRD (%) HRD 0 -10 Frequency (%) Frequency -1.0 0.0 1.0 1101001000 HRD (/100) Mean of Data Pair (ppm)
Mean HRD: 1.69 Median HRD: 0.81 Mean HRD: 1.69 Median HRD: 0.81 Precision: +/-10% Precision: +/-10%
T & H Precision Plot (Assay Pairs) T & H Precision Plot (Grouped Pairs) (U3O8Corp Sample DH U ME-MS61U Analysis) (U3O8Corp Sample DH U ME-MS61U Analysis)
1000 100 100 10 10 1 0.1 1
1101001000 (ppm) AD Median 10 100 Mean of Dat a Pair (ppm) Grouped Mean of Pair (ppm) Absolu te Difference (ppm)
10% 20% 30% 50% 10% 20% 30% 50%
Correlation Plot QQ Plot (U3O8Corp Sample DH U ME-MS61U Analysis) (U3O8Corp Sample DH U ME-MS61U Analysis)
1500 1500
1000 1000
500 500
0 0 0 200 400 600 800 1000 1200 1400 0 200 400 600 800 1000 1200 1400 Orig_Assay_U (ppm) Orig_A ssay_U (ppm) Duplicate_Assay_U (ppm) Duplicate_Assay_U (ppm)
P.CC= 1.00 S.CC= 0.99 Ref. Line y = 1.01x -1.65 Ref. Line y = 1.01x -1.65
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Appendix A – QAQC Summary Plots Page: 26 Berlin Project U3O8 Corp. Duplicates
Berlin Project Duplicate Analysis Feb 12 (Lab Pulp U ME-MS61U Analysis )
Orig_Assay Duplicate_A _U ssay_U Units Result No. Pairs: 24 24 Pearson CC: 1.00 Minimum: 5.60 4.30 ppm Spearman CC: 0.99 Max imum: 1,730.00 1,760.00 ppm Mean HA RD: 2.46 Mean: 231.33 230.47 ppm Median HARD: 2.02 Median 15.90 16.30 ppm Std. Deviation: 447.39 448.97 ppm Mean HRD: 0.16 Coefficient of Variation: 1.93 1.95 Median HRD -0.65
Mean vs. HARD Plot Rank HARD Plot (Lab Pulp U ME-MS61U Analysis ) (Lab Pulp U ME-MS61U Analysis )
15 100
10 95.83% of data are within 50 Precision limits 5 HARD (%) HARD 0 (%) HARD 1 10 100 1000 10000 0 0 102030405060708090100 Mean of Data Pair (ppm) Rank (%)
M ean HARD: 2.46 M edian HARD: 2. 02 Precision: 10% Precision: 10%
HRD Histogram Mean vs. HRD Plot (Lab Pulp U ME-MS61U Analysis ) (Lab Pulp U ME-MS61U Analysis )
80 20
60 10 40 20 0 HRD (%) HRD 0 -10 Frequency (%) Frequency -1.0 0.0 1.0 1 10 100 1000 10000 HRD (/100) Mean of Data Pair (ppm)
Mean HRD: 0.16 Median HRD: -0.65 Mean HRD: 0.16 Median HRD: -0.65 Precision: + /-10% Precision: + /-10%
T & H Precision Plot (Assay Pairs) T & H Precision Plot (Grouped Pairs) (Lab Pulp U ME-MS61U Analysis ) (Lab Pulp U ME-MS61U Analysis )
1000 1000 100 100 10 10 1 1 0.1 0.1
1 10 100 1000 10000 (ppm) AD Median 1 10 100 1000 Mean of Data Pair (ppm) Grouped Mean of Pair (ppm) Absolu te Differen (ppm) ce
10% 20% 30% 50% 10% 20% 30% 50%
Correlation Plot QQ Plot (Lab Pulp U ME-MS61U Analysis ) (Lab Pulp U ME-MS61U Analysis )
2000 2000 1500 1500 1000 1000 500 500 0 0 0 500 1000 1500 2000 0 500 1000 1500 2000 Orig_Assay_U (ppm) Orig_A ssay _U (ppm) Duplicate_Assay_U (ppm) Duplicate_Assay_U (ppm) Duplicate_Assay_U
P.CC= 1.00 S.CC= 0.99 Ref. Line y = 1.00x -1.56 Ref. Line y = 1.00x -1.56
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Appendix A – QAQC Summary Plots Page: 27 Berlin Project U3O8 Corp. Duplicates
Berlin Project Duplicate Analysis Feb 12 (U3O8Corp 10 Mesh U ME-MS81 Analysis)
Orig_Assay Duplicate_A _U ssay_U Units Result No. Pairs: 5 5 Pearson CC: 1.00 Minimum: 8.64 8.42 ppm Spearman CC: 1.00 Max imum: 788.00 838.00 ppm Mean HA RD: 1.91 Mean: 295.89 302.32 ppm Median HARD: 1.98 Median 258.00 248.00 ppm Std. Deviation: 272.99 292.17 ppm Mean HRD: 0.68 Coefficient of Variation: 0.92 0.97 Median HRD 1.29
Mean vs. HARD Plot Rank HARD Plot (U3O8Corp 10 Mesh U ME-MS81 Analysis) (U3O8Corp 10 Mesh U ME-MS81 Analysis)
10 100
5 50 HARD (%) HARD 0 (%) HARD 1 10 100 1000 0 0 102030405060708090100 Mean of Data Pair (ppm) Rank (%)
M ean HARD: 1.91 M edian HARD: 1. 98 Precision: 10% Precision: 10%
HRD Histogram Mean vs. HRD Plot (U3O8Corp 10 Mesh U ME-MS81 Analysis) (U3O8Corp 10 Mesh U ME-MS81 Analysis)
60 10
40 5 0 20 -5 HRD (%) HRD 0 -10 Frequency (%) Frequency -1.0 0.0 1.0 1 10 100 1000 HRD (/100) Mean of Data Pair (ppm)
M ean HRD: 0.68 M edian HRD: 1.29 M ean HRD: 0.68 M edian HRD: 1.29 Precision: + /-10% Precision: + /-10%
T & H Precision Plot (Assay Pairs) T & H Precision Plot (Grouped Pairs) (U3O8Corp 10 Mesh U ME-MS81 Analysis) (U3O8Corp 10 Mesh U ME-MS81 Analysis)
1000 100 10 1 There is not enough data to generate this plot. 0.1 1 10 100 1000 Mean of Data Pair (ppm) Absolute Difference Difference (ppm) Absolute
10% 20% 30% 50%
Correlation Plot QQ Plot (U3O8Corp 10 Mesh U ME-MS81 Analysis) (U3O8Corp 10 Mesh U ME-MS81 Analysis)
1000 1000
500 500 0 0 -50 0 0 200 400 600 800 1000 -50 0 0 200 400 600 800 1000 Orig_Assay_U (ppm)
Duplicate_Assay_U (ppm) Orig_Assay_U (ppm) Duplicate_Assay_U (ppm)
P.CC= 1.00 S.CC= 1.00 Ref. Line y = 1.07x -14.12 Ref. Line y = 1.07x -14.12
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Appendix A – QAQC Summary Plots Page: 28 Berlin Project U3O8 Corp. Duplicates
Berlin Project Duplicate Statistics Feb 12 (Trench Pulp Duplicate U ME-MS81 Analysis )
Orig_Assay Duplicate _A _U ssay_U Units Result No. Pairs: 12 12 Pearson CC: 1.00 Minimum: 41.90 41.80 ppm Spearman CC: 0.99 Max imum: 621.00 634.00 ppm Mean HA RD: 1.14 Mean: 219.20 217.23 ppm Median HA RD: 0.98 Median 170.50 169.50 ppm Std. Deviation: 170.46 168.95 ppm Mean HRD: 0.02 Coefficient of Variation: 0.78 0.78 Median HRD 0.21
Mean vs. HARD Plot Rank HARD Plot (Trench Pulp Duplicate U ME-MS81 Analysis ) (Trench Pulp Duplicate U ME-MS81 Analysis )
10 100
5 50 HARD (%) HARD 0 (%) HARD 11010010000 0 102030405060708090100 Mean of Data Pair (ppm) Rank (%)
Mean HARD: 1.14 Median HARD: 0. 98 Precision: 10% Precision: 10%
HRD Histogram Mean vs. HRD Plot (Trench Pulp Duplicate U ME-MS81 Analysis ) (Trench Pulp Duplicate U ME-MS81 Analysis )
100 10 5 50 0 -5 HRD (%) HRD 0 -10 Frequency (%) Frequency -1.0 0.0 1.0 1101001000 HRD (/100) Mean of Data Pair (ppm)
Mean HRD: 0.02 Median HRD: 0.21 Mean HRD: 0.02 Median HRD: 0.21 Precision: + /-10% Precision: + /-10%
T & H Precision Plot (Assay Pairs) T & H Precision Plot (Grouped Pairs) (Trench Pulp Duplicate U ME-MS81 Analysis ) (Trench Pulp Duplicate U ME-MS81 Analysis )
1000 100 100 10 10 1 0.1 1
10 100 1000 (ppm) AD Median 100 1000 Mean of Dat a Pair (ppm) Grouped Mean of Pair (ppm) Absolu te Difference (ppm)
10% 20% 30% 50% 10% 20% 30% 50%
Correlation Plot QQ Plot (Trench Pulp Duplicate U ME-MS81 Analysis ) (Trench Pulp Duplicate U ME-MS81 Analysis )
800 800 600 600 400 400 200 200 0 0 100 200 300 400 500 600 700 800 0 0 100 200 300 400 500 600 700 800 Orig_Assay_U (ppm)
Duplicate_Assay_U (ppm) Orig_A ssay_U (ppm) Duplicate_Assay_U (ppm)
P.CC= 1.00 S.CC= 0.99 Ref. Line y = 0.99x + 0. 24 Ref. Line y = 0.99x + 0. 24
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Appendix A – QAQC Summary Plots Page: 29 Berlin Project U3O8 Corp. Duplicates
Berlin Project Duplicate Analysis Feb 12 (U3O8Corp DH Duplicate U ME-MS81 Analysis)
Orig_Assay Duplicate_A _U ssay_U Units Result No. Pairs: 8 8 Pearson CC: 1.00 Minimum: 21.10 22.20 ppm Spearman CC: 1.00 Max imum: 730.00 793.00 ppm Mean HA RD: 2.04 Mean: 318.23 331.94 ppm Median HARD: 1.47 Median 285.75 287.50 ppm Std. Deviation: 256.80 271.59 ppm Mean HRD: -1.55 Coefficient of Variation: 0.81 0.82 Median HRD -0.67
Mean vs. HARD Plot Rank HARD Plot (U3O8Corp DH Duplicate U ME-MS81 Analysis) (U3O8Corp DH Duplicate U ME-MS81 Analysis)
10 100
5 50 HARD (%) HARD 0 (%) HARD 1 10 100 1000 0 0 102030405060708090100 Mean of Data Pair (ppm) Rank (%)
M ean HARD: 2.04 M edian HARD: 1. 47 Precision: 10% Precision: 10%
HRD Histogram Mean vs. HRD Plot (U3O8Corp DH Duplicate U ME-MS81 Analysis) (U3O8Corp DH Duplicate U ME-MS81 Analysis)
80 10 60 5 40 0 20 -5 HRD (%) HRD 0 -10 Frequency (%) Frequency -1.0 0.0 1.0 1 10 100 1000 HRD (/100) Mean of Data Pair (ppm)
Mean HRD: -1.55 Median HRD: -0.67 Mean HRD: -1.55 Median HRD: -0.67 Precision: + /-10% Precision: + /-10%
T & H Precision Plot (Assay Pairs) T & H Precision Plot (Grouped Pairs) (U3O8Corp DH Duplicate U ME-MS81 Analysis) (U3O8Corp DH Duplicate U ME-MS81 Analysis)
1000 100 10 1 There is not enough data to generate this plot. 0.1 10 100 1000 Mean of Data Pair (ppm) Absolu te Differen (ppm) ce
10% 20% 30% 50%
Correlation Plot QQ Plot (U3O8Corp DH Duplicate U ME-MS81 Analysis) (U3O8Corp DH Duplicate U ME-MS81 Analysis)
800 800 600 600 400 400 200 200 0 0 0 100 200 300 400 500 600 700 800 0 100 200 300 400 500 600 700 800 Orig_A ssay_U (ppm) Orig_A ssay_U (ppm) Duplicate_Assay_U (ppm) Duplicate_Assay_U (ppm) Duplicate_Assay_U
P.CC= 1.00 S.CC= 1.00 Ref. Line y = 1.05x -3.07 Ref. Line y = 1.05x -3.07
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Appendix A – QAQC Summary Plots Page: 30 Berlin Project U3O8 Corp. Duplicates
Berlin Project Duplicate Statistics Feb 12 (DH Pulp Duplicate U ME-MS81 Analysis)
Duplicate_A Assay_V_ ssay_V Units Result No. Pairs: 8 8 Pearson CC: 0.99 Minimum: 152.00 149.00 ppm Spearman CC: 0.98 Max imum: 3,320.00 3,410.00 ppm Mean HA RD: 2.25 Mean: 1,898.63 1,939.25 ppm Median HA RD: 0.77 Median 2,145.00 2,025.00 ppm Std. Deviation: 1,059.61 1,107.37 ppm Mean HRD: -0.47 Coefficient of Variation: 0.56 0.57 Median HRD 0.29
Mean vs. HARD Plot Rank HARD Plot (DH Pulp Duplicate U ME-MS81 Analysis) (DH Pulp Duplicate U ME-MS81 Analysis)
10 100
5 50 HARD (%) HARD 0 (%) HARD 1 10 100 1000 10000 0 0 102030405060708090100 Mean of Data Pair (ppm) Rank (%)
Mean HARD: 2.25 Median HARD: 0. 77 Precision: 10% Precision: 10%
HRD Histogram Mean vs. HRD Plot (DH Pulp Duplicate U ME-MS81 Analysis) (DH Pulp Duplicate U ME-MS81 Analysis)
80 10 60 5 40 0 20 -5 HRD (%) HRD 0 -10 Frequency (%) Frequency -1.0 0.0 1.0 1 10 100 1000 10000 HRD (/100) Mean of Data Pair (ppm)
Mean HRD: -0.47 Median HRD: 0.29 Mean HRD: -0.47 Median HRD: 0.29 Precision: + /-10% Precision: + /-10%
T & H Precision Plot (Assay Pairs) T & H Precision Plot (Grouped Pairs) (DH Pulp Duplicate U ME-MS81 Analysis) (DH Pulp Duplicate U ME-MS81 Analysis)
10000 1000 100 10 There is not enough data to generate this plot. 1 100 1000 10000 Mean of Data Pair (ppm) Absolute Differen (ppm) ce
10% 20% 30% 50%
Correlation Plot QQ Plot (DH Pulp Duplicate U ME-MS81 Analysis) (DH Pulp Duplicate U ME-MS81 Analysis)
4000 4000 3000 3000 2000 2000 1000 1000 0 0 1000 2000 3000 4000 0 0 1000 2000 3000 4000 Assay_V_ (ppm)
Duplicate_Assay_V (ppm) Duplicate_Assay_V Assay_V_ (ppm) Duplicate_Assay_V (ppm) Duplicate_Assay_V
P.CC= 0.99 S.CC= 0.98 Ref. Line y = 1.03x -15.30 Ref. Line y = 1.04x -28.23
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Appendix A – QAQC Summary Plots Page: 31 Berlin Project U3O8 Corp. Duplicates
Berlin Project Duplicate Statistics Feb 12 (Lab Pulp Duplicate U ME-MS81 Analysis )
Orig_Assay Duplicate_A _U ssay_U Units Result No. Pairs: 7 7 Pearson CC: 1.00 Minimum: 16.60 16.00 ppm Spearman CC: 1.00 Max imum: 953.00 947.00 ppm Mean HA RD: 1.30 Mean: 377.69 378.21 ppm Median HA RD: 1.33 Median 219.00 226.00 ppm Std. Deviation: 371.69 373.58 ppm Mean HRD: 0.47 Coefficient of Variation: 0.98 0.99 Median HRD 1.04
Mean vs. HARD Plot Rank HARD Plot (Lab Pulp Duplicate U ME-MS81 Analysis ) (Lab Pulp Duplicate U ME-MS81 Analysis )
10 100
5 50 HARD (%) HARD 0 (%) HARD 11010010000 0 102030405060708090100 Mean of Data Pair (ppm) Rank (%)
Mean HARD: 1.30 Median HARD: 1. 33 Precision: 10% Precision: 10%
HRD Histogram Mean vs. HRD Plot (Lab Pulp Duplicate U ME-MS81 Analysis ) (Lab Pulp Duplicate U ME-MS81 Analysis )
100 10 5 50 0 -5 HRD (%) HRD 0 -10 Frequency (%) Frequency -1.0 0.0 1.0 1101001000 HRD (/100) Mean of Data Pair (ppm)
Mean HRD: 0.47 Median HRD: 1.04 Mean HRD: 0.47 Median HRD: 1.04 Precision: + /-10% Precision: + /-10%
T & H Precision Plot (Assay Pairs) T & H Precision Plot (Grouped Pairs) (Lab Pulp Duplicate U ME-MS81 Analysis ) (Lab Pulp Duplicate U ME-MS81 Analysis )
1000 100 10 1 There is not enough data to generate this plot. 0.1 10 100 1000 Mean of Dat a Pair (ppm) Absolute Differen (ppm) ce
10% 20% 30% 50%
Correlation Plot QQ Plot (Lab Pulp Duplicate U ME-MS81 Analysis ) (Lab Pulp Duplicate U ME-MS81 Analysis )
1000 1000
500 500
0 0 0 200 400 600 800 1000 0 200 400 600 800 1000 Orig_A ssay_U (ppm) Orig_A ssay_U (ppm) Duplicate_Assay_U (ppm) Duplicate_Assay_U (ppm) Duplicate_Assay_U
P.CC= 1.00 S.CC= 1.00 Ref. Line y = 1.00x -1.21 Ref. Line y = 1.00x -1.21
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Appendix A – QAQC Summary Plots Page: 32 Berlin Project U3O8 Corp. Duplicates
Berlin Project Duplicate Statistics Feb 12 (Lab Pulp Duplicate P ME-MS81 Analysis )
Duplicate_A Assay_P_ ssay_P Units Result No. Pairs: 10 10 Pearson CC: 1.00 Minimum: 0.00 0.00 ppm Spearman CC: 0.96 Max imum: 3,740.00 3,670.00 ppm Mean HA RD: 2.15 Mean: 782.80 764.20 ppm Median HARD: 0.64 Median 155.50 135.00 ppm Std. Deviation: 1,297.20 1,280.84 ppm Mean HRD: 1.87 Coefficient of Variation: 1.66 1.68 Median HRD 0.17
Mean vs. HARD Plot Rank HARD Plot (Lab Pulp Duplicate P ME-MS81 Analysis ) (Lab Pulp Duplicate P ME-MS81 Analysis )
10 100
5 50 HARD (%) HARD 0 (%) HARD 1 10 100 1000 10000 0 0 102030405060708090100 Mean of Data Pair (ppm) Rank (%)
M ean HARD: 2.15 M edian HARD: 0. 64 Precision: 10% Precision: 10%
HRD Histogram Mean vs. HRD Plot (Lab Pulp Duplicate P ME-MS81 Analysis ) (Lab Pulp Duplicate P ME-MS81 Analysis )
80 10 60 5 40 0 20 -5 HRD (%) HRD 0 -10 Frequency (%) Frequency -1.0 0.0 1.0 1 10 100 1000 10000 HRD (/100) Mean of Data Pair (ppm)
M ean HRD: 1.87 M edian HRD: 0.17 M ean HRD: 1.87 M edian HRD: 0.17 Precision: + /-10% Precision: + /-10%
T & H Precision Plot (Assay Pairs) T & H Precision Plot (Grouped Pairs) (Lab Pulp Duplicate P ME-MS81 Analysis ) (Lab Pulp Duplicate P ME-MS81 Analysis )
10000 1000 100 10 There is not enough data to generate this plot. 1 10 100 1000 10000 Mean of Data Pair (ppm) Absolu te Difference (ppm)
10% 20% 30% 50%
Correlation Plot QQ Plot (Lab Pulp Duplicate P ME-MS81 Analysis ) (Lab Pulp Duplicate P ME-MS81 Analysis )
4000 4000
2000 2000
0 0
-20 00 -20 00 0 1000 2000 3000 4000 0 1000 2000 3000 4000 Assay_P_ (ppm) Assay_P_ (ppm) Duplicate_Assay_P (ppm) Duplicate_Assay_P (ppm)
P.CC= 1.00 S.CC= 0.96 Ref. Line y = 0.99x -8.64 Ref. Line y = 0.99x -8.64
Printed: 12-Feb-2012 21:32:33 Data Imported: 04-Feb-2012 21:45:53 Page 1
Appendix A – QAQC Summary Plots Page: 33 Berlin Project U3O8 Corp. Duplicates
Berlin Project Duplicate Statistics Feb 12 (Trench Pulp Duplicate V ME-MS81 Analysis )
Duplicate_A Assay_V_ ssay_V Units Result No. Pairs: 12 12 Pearson CC: 1.00 Minimum: 284.00 297.00 ppm Spearman CC: 1.00 Max imum: 10,000.00 10,000.00 ppm Mean HA RD: 0.96 Mean: 2,237.17 2,230.58 ppm Median HARD: 0.94 Median 1,460.00 1,435.00 ppm Std. Deviation: 2,733.69 2,731.47 ppm Mean HRD: -0.05 Coefficient of Variation: 1.22 1.22 Median HRD 0.29
Mean vs. HARD Plot Rank HARD Plot (Trench Pulp Duplicate V ME-MS81 Analysis )(Trench Pulp Duplicate V ME-MS81 Analysis )
10 100
5 50 HARD (%) HARD 0 (%) HARD 1 10 100 1000 10000 0 0 102030405060708090100 Mean of Data Pair (ppm) Rank (%)
M ean HARD: 0.96 M edian HARD: 0. 94 Precision: 10% Precision: 10%
HRD Histogram Mean vs. HRD Plot (Trench Pulp Duplicate V ME-MS81 Analysis )(Trench Pulp Duplicate V ME-MS81 Analysis )
100 10 5 50 0 -5 HRD (%) HRD 0 -10 Frequency (%) Frequency -1.0 0.0 1.0 1 10 100 1000 10000 HRD (/100) Mean of Data Pair (ppm)
Mean HRD: -0.05 M edian HRD: 0.29 Mean HRD: -0.05 M edian HRD: 0.29 Precision: + /-10% Precision: + /-10%
T & H Precision Plot (Assay Pairs) T & H Precision Plot (Grouped Pairs) (Trench Pulp Duplicate V ME-MS81 Analysis )(Trench Pulp Duplicate V ME-MS81 Analysis )
10000 1000 1000 100 100 10 1 10
100 1000 10000 (ppm) AD Median 1000 10000 Mean of Data Pair (ppm) Grouped Mean of Pair (ppm) Absolu te Difference (ppm)
10% 20% 30% 50% 10% 20% 30% 50%
Correlation Plot QQ Plot (Trench Pulp Duplicate V ME-MS81 Analysis )(Trench Pulp Duplicate V ME-MS81 Analysis )
10000 10000
5000 5000
0 0 0 2000 4000 6000 8000 10000 0 2000 4000 6000 8000 10000 Assay_V_ (ppm) Assay_V_ (ppm) Duplicate_Assay_V (ppm) Duplicate_Assay_V (ppm)
P.CC= 1.00 S.CC= 1.00 Ref. Line y = 1.00x -4.69 Ref. Line y = 1.00x -4.69
Printed: 12-Feb-2012 21:39:55 Data Imported: 04-Feb-2012 21:45:53 Page 1
Appendix A – QAQC Summary Plots Page: 34 Berlin Project U3O8 Corp. Duplicates
Berlin Project Duplicate Statistics Feb 12 (DH Pulp Duplicate V ME-MS81 Analysis )
Duplicate_A Assay_V_ ssay_V Units Result No. Pairs: 8 8 Pearson CC: 0.99 Minimum: 152.00 149.00 ppm Spearman CC: 0.98 Max imum: 3,320.00 3,410.00 ppm Mean HA RD: 2.25 Mean: 1,898.63 1,939.25 ppm Median HARD: 0.77 Median 2,145.00 2,025.00 ppm Std. Deviation: 1,059.61 1,107.37 ppm Mean HRD: -0.47 Coefficient of Variation: 0.56 0.57 Median HRD 0.29
Mean vs. HARD Plot Rank HARD Plot (DH Pulp Duplicate V ME-MS81 Analysis ) (DH Pulp Duplicate V ME-MS81 Analysis )
10 100
5 50 HARD (%) HARD 0 (%) HARD 1 10 100 1000 10000 0 0 102030405060708090100 Mean of Data Pair (ppm) Rank (%)
M ean HARD: 2.25 M edian HARD: 0. 77 Precision: 10% Precision: 10%
HRD Histogram Mean vs. HRD Plot (DH Pulp Duplicate V ME-MS81 Analysis ) (DH Pulp Duplicate V ME-MS81 Analysis )
80 10 60 5 40 0 20 -5 HRD (%) HRD 0 -10 Frequency (%) Frequency -1.0 0.0 1.0 1 10 100 1000 10000 HRD (/100) Mean of Data Pair (ppm)
Mean HRD: -0.47 M edian HRD: 0.29 Mean HRD: -0.47 M edian HRD: 0.29 Precision: + /-10% Precision: + /-10%
T & H Precision Plot (Assay Pairs) T & H Precision Plot (Grouped Pairs) (DH Pulp Duplicate V ME-MS81 Analysis ) (DH Pulp Duplicate V ME-MS81 Analysis )
10000 1000 100 10 There is not enough data to generate this plot. 1 100 1000 10000 Mean of Data Pair (ppm) Absolu te Differen (ppm) ce
10% 20% 30% 50%
Correlation Plot QQ Plot (DH Pulp Duplicate V ME-MS81 Analysis ) (DH Pulp Duplicate V ME-MS81 Analysis )
4000 4000 3000 3000 2000 2000 1000 1000 0 0 1000 2000 3000 4000 0 0 1000 2000 3000 4000 Assay_V_ (ppm)
Duplicate_Assay_V (ppm) Duplicate_Assay_V Assay_V_ (ppm) Duplicate_Assay_V (ppm) Duplicate_Assay_V
P.CC= 0.99 S.CC= 0.98 Ref. Line y = 1.03x -15.30 Ref. Line y = 1.04x -28.23
Printed: 12-Feb-2012 20:25:39 Data Imported: 04-Feb-2012 21:45:53 Page 1
Appendix A – QAQC Summary Plots Page: 35 Berlin Project U3O8 Corp. Duplicates
Berlin Project Duplicate Statistics Feb 12 (Lab Pulp Duplicate V ME-MS81 Analysis )
Duplicate_A Assay_V_ ssay_V Units Result No. Pairs: 10 10 Pearson CC: 0.99 Minimum: 1,260.00 1,280.00 ppm Spearman CC: 0.96 Max imum: 5,980.00 5,610.00 ppm Mean HA RD: 1.99 Mean: 2,889.00 2,876.00 ppm Median HA RD: 1.19 Median 2,310.00 2,325.00 ppm Std. Deviation: 1,320.31 1,277.94 ppm Mean HRD: 0.21 Coefficient of Variation: 0.46 0.44 Median HRD 0.57
Mean vs. HARD Plot Rank HARD Plot (Lab Pulp Duplicate V ME-MS81 Analysis ) (Lab Pulp Duplicate V ME-MS81 Analysis )
10 100
5 50 HARD (%) HARD 0 (%) HARD 1 10 100 1000 10000 0 0 102030405060708090100 Mean of Data Pair (ppm) Rank (%)
Mean HARD: 1.99 Median HARD: 1. 19 Precision: 10% Precision: 10%
HRD Histogram Mean vs. HRD Plot (Lab Pulp Duplicate V ME-MS81 Analysis ) (Lab Pulp Duplicate V ME-MS81 Analysis )
80 10 60 5 40 0 20 -5 HRD (%) HRD 0 -10 Frequency (%) Frequency -1.0 0.0 1.0 1 10 100 1000 10000 HRD (/100) Mean of Data Pair (ppm)
Mean HRD: 0.21 Median HRD: 0.57 Mean HRD: 0.21 Median HRD: 0.57 Precision: + /-10% Precision: + /-10%
T & H Precision Plot (Assay Pairs) T & H Precision Plot (Grouped Pairs) (Lab Pulp Duplicate V ME-MS81 Analysis ) (Lab Pulp Duplicate V ME-MS81 Analysis )
10000
1000
100 There is not enough data to generate this plot. 10 1000 10000 Mean of Data Pair (ppm) Absolute Differen (ppm) ce
10% 20% 30% 50%
Correlation Plot QQ Plot (Lab Pulp Duplicate V ME-MS81 Analysis ) (Lab Pulp Duplicate V ME-MS81 Analysis )
6000 6000
4000 4000 2000 2000 0 0 1000 2000 3000 4000 5000 6000 0 0 1000 2000 3000 4000 5000 6000 Assay_V_ (ppm)
Duplicate_Assay_V (ppm) Duplicate_Assay_V Assay_V_ (ppm) Duplicate_Assay_V (ppm) Duplicate_Assay_V
P.CC= 0.99 S.CC= 0.96 Ref. Line y = 0.96x + 106. 86 Ref. Line y = 0.96x + 99. 17
Printed: 12-Feb-2012 21:25:34 Data Imported: 04-Feb-2012 21:45:53 Page 1
Appendix A – QAQC Summary Plots Page: 36 Berlin Project U3O8 Corp. Duplicates
Berlin Project Duplicate Statistics Feb 12 (Trench Pulp Duplicate Y ME-MS81 Analysis )
Duplicate_A Assay_Y_ ssay_Y Units Result No. Pairs: 12 12 Pearson CC: 1.00 Minimum: 33.80 33.60 ppm Spearman CC: 1.00 Max imum: 601.00 623.00 ppm Mean HA RD: 0.94 Mean: 221.64 222.18 ppm Median HA RD: 0.72 Median 126.00 126.00 ppm Std. Deviation: 193.84 198.00 ppm Mean HRD: 0.21 Coefficient of Variation: 0.87 0.89 Median HRD 0.44
Mean vs. HARD Plot Rank HARD Plot (Trench Pulp Duplicate Y ME-MS81 Analysis )(Trench Pulp Duplicate Y ME-MS81 Analysis )
10 100
5 50 HARD (%) HARD 0 (%) HARD 11010010000 0 102030405060708090100 Mean of Data Pair (ppm) Rank (%)
Mean HARD: 0.94 Median HARD: 0.72 Precision: 10% Precision: 10%
HRD Histogram Mean vs. HRD Plot (Trench Pulp Duplicate Y ME-MS81 Analysis )(Trench Pulp Duplicate Y ME-MS81 Analysis )
100 10 5 50 0 -5 HRD (%) HRD 0 -10 Frequency (%) Frequency -1.0 0.0 1.0 1101001000 HRD (/100) Mean of Data Pair (ppm)
Mean HRD: 0.21 Median HRD: 0.44 Mean HRD: 0.21 Median HRD: 0.44 Precision: + /-10% Precision: +/-10%
T & H Precision Plot (Assay Pairs) T & H Precision Plot (Grouped Pairs) (Trench Pulp Duplicate Y ME-MS81 Analysis )(Trench Pulp Duplicate Y ME-MS81 Analysis )
1000 100 100 10 10 1 0.1 1
10 100 1000 (ppm) AD Median 100 1000 Mean of Dat a Pair (ppm) Grouped Mean of Pair (ppm) Absolute Difference (ppm) Absolute
10% 20% 30% 50% 10% 20% 30% 50%
Correlation Plot QQ Plot (Trench Pulp Duplicate Y ME-MS81 Analysis )(Trench Pulp Duplicate Y ME-MS81 Analysis )
800 800 600 600 400 400 200 200 0 0 0 100 200 300 400 500 600 700 800 0 100 200 300 400 500 600 700 800 Assay_Y_ (ppm) Assay_Y_ (ppm) Duplicate_Assay_Y (ppm) Duplicate_Assay_Y (ppm)
P.CC= 1.00 S.CC= 1.00 Ref. Line y = 1.02x -4.11 Ref. Line y = 1.02x -4.11
Printed: 12-Feb-2012 21:42:08 Data Imported: 04-Feb-2012 21:45:53 Page 1
Appendix A – QAQC Summary Plots Page: 37 Berlin Project U3O8 Corp. Duplicates
Berlin Project Duplicate Statistics Feb 12 (Lab Pulp Duplicate Y ME-MS81 Analysis )
Duplicate_A Assay_Y_ ssay_Y Units Result No. Pairs: 10 10 Pearson CC: 1.00 Minimum: 14.30 14.10 ppm Spearman CC: 1.00 Max imum: 831.00 837.00 ppm Mean HA RD: 0.98 Mean: 342.39 343.45 ppm Median HA RD: 0.87 Median 309.50 300.00 ppm Std. Deviation: 299.72 302.94 ppm Mean HRD: 0.33 Coefficient of Variation: 0.88 0.88 Median HRD -0.11
Mean vs. HARD Plot Rank HARD Plot (Lab Pulp Duplicate Y ME-MS81 Analysis ) (Lab Pulp Duplicate Y ME-MS81 Analysis )
10 100
5 50 HARD (%) HARD 0 (%) HARD 11010010000 0 102030405060708090100 Mean of Data Pair (ppm) Rank (%)
Mean HARD: 0.98 Median HARD: 0.87 Precision: 10% Precision: 10%
HRD Histogram Mean vs. HRD Plot (Lab Pulp Duplicate Y ME-MS81 Analysis ) (Lab Pulp Duplicate Y ME-MS81 Analysis )
100 10 5 50 0 -5 HRD (%) HRD 0 -10 Frequency (%) Frequency -1.0 0.0 1.0 1101001000 HRD (/100) Mean of Data Pair (ppm)
Mean HRD: 0.33 Median HRD: -0.11 M ean HRD: 0.33 Median HRD: -0.11 Precision: + /-10% Precision: + /-10%
T & H Precision Plot (Assay Pairs) T & H Precision Plot (Grouped Pairs) (Lab Pulp Duplicate Y ME-MS81 Analysis ) (Lab Pulp Duplicate Y ME-MS81 Analysis )
1000 100 10 1 There is not enough data to generate this plot. 0.1 10 100 1000 Mean of Dat a Pair (ppm) Absolu te Difference (ppm)
10% 20% 30% 50%
Correlation Plot QQ Plot (Lab Pulp Duplicate Y ME-MS81 Analysis ) (Lab Pulp Duplicate Y ME-MS81 Analysis )
1000 1000
500 500
0 0 0 200 400 600 800 1000 0 200 400 600 800 1000 Assay_Y_ (ppm) Assay_Y_ (ppm) Duplicate_Assay_Y (ppm) Duplicate_Assay_Y (ppm)
P.CC= 1.00 S.CC= 1.00 Ref. Line y = 1.01x -2.57 Ref. Line y = 1.01x -2.57
Printed: 12-Feb-2012 21:29:22 Data Imported: 04-Feb-2012 21:45:53 Page 1
Appendix A – QAQC Summary Plots Page: 38 Berlin Project U3O8 Corp. Duplicates
Berlin Project Duplicate Statistics Feb 12 (Trench Pulp Duplicate Ni ME-MS81 Analysis )
Duplicate_A Assay_Ni_ ssay_Ni Units Result No. Pairs: 12 12 Pearson CC: 1.00 Minimum: 0.00 0.00 ppm Spearman CC: 1.00 Max imum: 344.00 369.00 ppm Mean HA RD: 4.34 Mean: 90.92 94.17 ppm Median HA RD: 2.30 Median 47.00 42.00 ppm Std. Deviation: 103.84 110.59 ppm Mean HRD: -1.48 Coefficient of Variation: 1.14 1.17 Median HRD -0.79
Mean vs. HARD Plot Rank HARD Plot (Trench Pulp Duplicate Ni ME-MS81 Analysis (Trench Pulp Duplicate Ni ME-MS81 Analysis
20 100 15 83.33% of data are within 10 50 Precision limits 5 HARD (%) HARD 0 (%) HARD 11010010000 0 102030405060708090100 Mean of Data Pair (ppm) Rank (%)
Mean HARD: 4.34 Median HARD: 2. 30 Precision: 10% Precision: 10%
HRD Histogram Mean vs. HRD Plot (Trench Pulp Duplicate Ni ME-MS81 Analysis (Trench Pulp Duplicate Ni ME-MS81 Analysis
60 20
40 10 0 20 -10 HRD (%) HRD 0 -20 Frequency (%) Frequency -1.0 0.0 1.0 1101001000 HRD (/100) Mean of Data Pair (ppm)
Mean HRD: -1.48 Median HRD: -0.79 Mean HRD: -1.48 Median HRD: -0.79 Precision: + /-10% Precision: + /-10%
T & H Precision Plot (Assay Pairs) T & H Precision Plot (Grouped Pairs) (Trench Pulp Duplicate Ni ME-MS81 Analysis (Trench Pulp Duplicate Ni ME-MS81 Analysis
1000 100 100 10 10 1 0.1 1
1101001000 (ppm) AD Median 10 100 Mean of Dat a Pair (ppm) Grouped Mean of Pair (ppm) Absolu te Differen (ppm) ce
10% 20% 30% 50% 10% 20% 30% 50%
Correlation Plot QQ Plot (Trench Pulp Duplicate Ni ME-MS81 Analysis (Trench Pulp Duplicate Ni ME-MS81 Analysis
400 400
200 200
0 0
-20 0 -20 0 0100200300400 0 100 200 300 400 Assay_Ni_ (ppm) Assay_Ni_ (ppm) Duplicate_Assay_Ni (ppm) Duplicate_Assay_Ni (ppm)
P.CC= 1.00 S.CC= 1.00 Ref. Line y = 1.06x -2.56 Ref. Line y = 1.06x -2.56
Printed: 12-Feb-2012 21:38:32 Data Imported: 04-Feb-2012 21:45:53 Page 1
Appendix A – QAQC Summary Plots Page: 39 Berlin Project U3O8 Corp. Duplicates
Berlin Project Duplicate Statistics Feb 12 (Lab Pulp Duplicate Ni ME-MS81 Analysis)
Duplicate_A Assay_Ni_ ssay_Ni Units Result No. Pairs: 10 10 Pearson CC: 1.00 Minimum: 2.10 2.30 ppm Spearman CC: 1.00 Max imum: 183.50 186.00 ppm Mean HA RD: 1.91 Mean: 87.40 87.95 ppm Median HARD: 1.42 Median 95.70 91.60 ppm Std. Deviation: 73.41 74.61 ppm Mean HRD: -0.11 Coefficient of Variation: 0.84 0.85 Median HRD -0.76
Mean vs. HARD Plot Rank HARD Plot (Lab Pulp Duplicate Ni ME-MS81 Analysis) (Lab Pulp Duplicate Ni ME-MS81 Analysis)
10 100
5 50 HARD (%) HARD 0 (%) HARD 1 10 100 1000 0 0 102030405060708090100 Mean of Data Pair (ppm) Rank (%)
M ean HARD: 1.91 M edian HARD: 1. 42 Precision: 10% Precision: 10%
HRD Histogram Mean vs. HRD Plot (Lab Pulp Duplicate Ni ME-MS81 Analysis) (Lab Pulp Duplicate Ni ME-MS81 Analysis)
80 10 60 5 40 0 20 -5 HRD (%) HRD 0 -10 Frequency (%) Frequency -1.0 0.0 1.0 1101001000 HRD (/100) Mean of Data Pair (ppm)
Mean HRD: -0.11 Median HRD: -0.76 Mean HRD: -0.11 Median HRD: -0.76 Precision: + /-10% Precision: + /-10%
T & H Precision Plot (Assay Pairs) T & H Precision Plot (Grouped Pairs) (Lab Pulp Duplicate Ni ME-MS81 Analysis) (Lab Pulp Duplicate Ni ME-MS81 Analysis)
100
10
1 There is not enough data to generate this plot. 0.1 1 10 100 1000 Mean of Data Pair (ppm) Absolute Difference Difference (ppm) Absolute
10% 20% 30% 50%
Correlation Plot QQ Plot (Lab Pulp Duplicate Ni ME-MS81 Analysis) (Lab Pulp Duplicate Ni ME-MS81 Analysis)
200 200 150 150 100 100 50 50 0 0 0 20 40 60 80 100 120 140 160 180 200 0 20 40 60 80 100 120 140 160 180 200 Assay_Ni_ (ppm) Assay_Ni_ (ppm) Duplicate_Assay_Ni (ppm) Duplicate_Assay_Ni (ppm)
P.CC= 1.00 S.CC= 1.00 Ref. Line y = 1.02x -0.82 Ref. Line y = 1.02x -0.82
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Appendix A – QAQC Summary Plots Page: 40 Berlin Project U3O8 Corp. Duplicates
Berlin Project Duplicate Statistics Feb 12 (Trench Pulp Duplicate Nd ME-MS81 Analysis )
Duplicate_A Assay_Nd_ ssay_Nd Units Result No. Pairs: 12 12 Pearson CC: 1.00 Minimum: 9.10 8.90 ppm Spearman CC: 1.00 Max imum: 195.50 196.50 ppm Mean HA RD: 0.98 Mean: 76.83 77.26 ppm Median HARD: 0.85 Median 47.60 49.10 ppm Std. Deviation: 65.08 64.60 ppm Mean HRD: -0.54 Coefficient of Variation: 0.85 0.84 Median HRD -0.32
Mean vs. HARD Plot Rank HARD Plot (Trench Pulp Duplicate Nd ME-MS81 Analysis (Trench Pulp Duplicate Nd ME-MS81 Analysis
10 100
5 50 HARD (%) HARD 0 (%) HARD 1 10 100 1000 0 0 102030405060708090100 Mean of Data Pair (ppm) Rank (%)
M ean HARD: 0.98 M edian HARD: 0. 85 Precision: 10% Precision: 10%
HRD Histogram Mean vs. HRD Plot (Trench Pulp Duplicate Nd ME-MS81 Analysis (Trench Pulp Duplicate Nd ME-MS81 Analysis
100 10 5 50 0 -5 HRD (%) HRD 0 -10 Frequency (%) Frequency -1.0 0.0 1.0 1 10 100 1000 HRD (/100) Mean of Data Pair (ppm)
Mean HRD: -0.54 Median HRD: -0.32 Mean HRD: -0.54 Median HRD: -0.32 Precision: + /-10% Precision: + /-10%
T & H Precision Plot (Assay Pairs) T & H Precision Plot (Grouped Pairs) (Trench Pulp Duplicate Nd ME-MS81 Analysis (Trench Pulp Duplicate Nd ME-MS81 Analysis
100 100
10 10
1 1
0.1 0.1
1 10 100 1000 (ppm) AD Median 10 100 Mean of Data Pair (ppm) Grouped Mean of Pair (ppm) Absolu te Difference (ppm)
10% 20% 30% 50% 10% 20% 30% 50%
Correlation Plot QQ Plot (Trench Pulp Duplicate Nd ME-MS81 Analysis (Trench Pulp Duplicate Nd ME-MS81 Analysis
200 200 150 150 100 100 50 50 0 0 20 40 60 80 100 120 140 160 180 200 0 0 20 40 60 80 100 120 140 160 180 200 Assay_Nd_ (ppm) Assay_Nd_ (ppm) Duplicate_Assay_Nd (ppm) Duplicate_Assay_Nd (ppm) P.CC= 1.00 S.CC= 1.00 Ref. Line y = 0.99x + 1. 02 Ref. Line y = 0.99x + 1. 02
Printed: 12-Feb-2012 21:43:26 Data Imported: 04-Feb-2012 21:45:53 Page 1
Appendix A – QAQC Summary Plots Page: 41
Appendix B
Qualified Persons Certificates
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
Appendix B – Qualified Persons Certificates Page: 1
Certificate of Qualified Person
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.
4. I have practiced my profession continuously since 1963.
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 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.
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 parts of the Study for which I am responsible have 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 Toronto, Ontario, Canada on 2nd March, 2012.
[signed] G.R. Goode Metallurgical Consultant P.Eng, MCIMM, FAusIMM
Appendix B – Qualified Persons Certificates Page: 2
Certificate of Qualified Person
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.
4. I have practiced my profession continuously since 1982.
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 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.
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 Parts of the Study for which I am responsible for have 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 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
Appendix B – Qualified Persons Certificates Page: 3