INTEGRATED MINING AND PRECONCENTRATION SYSTEMS FOR NICKEL SULFIDE ORES
by
TRENT WILLIAM WEATHERWAX
BS Michigan Technological University, 2003
THESIS SUBMITTED IN PARTIAL FULFILMENT OF REQUIREMENTS FOR THE DEGREE OF
MASTER OF APPLIED SCIENCE
in
THE FACULTY OF GRADUATE STUDIES
(Mining Engineering)
THE UNIVERSITY OF BRITISH COLUMBIA
December, 2007
© Trent William Weatherwax, 2007 Abstract
As part of a strategic research initiative at UBC to design and evaluate integrated underground mining and mineral processing systems, work has been done to determine how to utilize the coarse rejects of pre concentration in the underground environment. An amenability study for nine orebodies from four of Xstrata Nickel’s Ontario operations evaluated both processing and waste disposal methods. Metallurgically the orebodies showed amenability to dense media separation and conductivity sorting. The dense media results showed high mass rejections and high metal recoveries for all nine orebodies. Conductivity sorter results were not as consistent, but still showed good results. Dense media rejects were examined to determine the applicability of their use in rockfills and composite minefills. The geotechnical properties indicated that the rejects would provide a competent material for minefills. The mix designs were examined for both strength and rheological properties and showed that fills utilizing rejects were comparable to fills currently used by industry. Composite fills containing rejects had significantly lower void ratios, decreasing cement requirements for a given strength requirement. Conceptual designs for pre concentration systems based on the metallurgical, reject characterization, and mix design were developed for each of the four mines in the study. The designs took into consideration the current mining plans.
ii Table of Contents
Abstract ...... ii Table of Contents ...... iii List of Tables...... vi List of Figures ...... vii Chapter 1 Introduction...... 1 1.1 Underground Preconcentration ...... 1 1.2 Objectives...... 2 1.3 The Mines and Ores of Xstrata Nickel...... 2 1.4 Methodology ...... 4 Chapter 2 Experimental Procedure...... 6 2.1 Introduction...... 6 2.2 Sample Collecting and Classification ...... 6 2.2.1 Sample Collecting...... 6 2.2.1 Mineralogical and Physical Characterization...... 6 2.3 Metallurgical Preconcentration Testing ...... 7 2.3.1 Size Assay...... 7 2.3.2 Feed Preparation...... 7 2.3.3 Dense Media Separation...... 7 2.3.4 Conductivity Sorting...... 9 2.4 Geotechnical Classification of Rejects...... 10 2.4.1 Particle Size Distribution ...... 10 2.4.2 Particle Shape...... 11 2.4.3 Adsorption and Specific Gravity...... 11 2.4.4 Void Space ...... 11 2.4.5 Hardness (Strength)...... 12 2.4.6 Chemical Composition...... 12 2.5 Backfill Mix Evaluation...... 12 2.5.1 Fill Mixing...... 12 2.5.2 Slump Testing ...... 13 2.5.3 UCS Cylinders and Testing...... 14 Chapter 3 Metallurgical Results ...... 15 3.1 Introduction...... 15 3.2 Literature Review...... 15 3.2.1 Preconcentration Plants...... 15 3.2.1.1 Dense Media Separation...... 16 3.2.1.2 Ore Sorting...... 17 3.2.2 Metallurgical Results from Preconcentration...... 19 3.3 Procedures...... 21 3.4 Results and Discussion...... 21 3.4.1 Mineralogical and Physical Characterization...... 21 3.4.2 Size Assay...... 23 3.4.3 Dense Media Separation...... 28 3.4.3.1 Overall Preconcentration...... 28 3.4.3.2 Precious Metal Recovery ...... 30
iii 3.4.3.3 MgO Rejection...... 30 3.4.4 Conductivity Sorting...... 32 3.4.4.1 Overall Preconcentration...... 32 3.4.4.2 MgO Rejection...... 34 3.5 Conclusions...... 34 3.6 Recommendations...... 35 Chapter 4 Classification of Preconcentration Rejects...... 36 4.1 Introduction...... 36 4.2 Literature Review...... 36 4.2.1 Particle Size Distribution ...... 37 4.2.2 Particle Shape...... 39 4.2.3 Void Ratio...... 40 4.2.4 Adsorption and Porosity...... 40 4.2.5 Specific Gravity...... 41 4.2.6 Hardness and Durability...... 41 4.2.7 Particle Strength...... 42 4.2.8 Chemical Composition...... 43 4.2.9 How to Compare Minefill Aggregate to Traditional Aggregate...... 43 4.3 Procedures...... 44 4.4 Results and Discussion...... 44 4.4.1 Geotechnical Properties...... 44 4.4.1.1 Particle size distribution...... 47 4.4.1.2 Particle shape...... 49 4.4.1.3 Void Ratio ...... 51 4.4.1.4 Adsorption and Porosity...... 52 4.4.1.5 Specific Gravity...... 53 4.4.1.6 Hardness and Durability...... 53 4.4.1.7 Strength ...... 53 4.4.1.8 Chemical Composition...... 54 4.5 Conclusions and Recommendations...... 55 Chapter 5 Testing of Mix Designs...... 56 5.1 Introduction...... 56 5.2 Literature Review...... 56 5.2.1 Current Backfill Practice and Performance...... 56 5.2.2 Composite Fills ...... 58 5.2.3 Design of Composite fills...... 61 5.2.4 Rheological Estimation...... 64 5.3 Procedures...... 65 5.4 Results and Discussion...... 66 5.4.1 Reject Based Rockfills...... 66 5.4.1.1 Physical Characteristics...... 67 5.4.1.2 UCS Test Results ...... 68 5.4.2 Composite Fills ...... 78 5.4.2.1 Physical Characteristics...... 81 5.4.2.2 Slump Test Results...... 85 5.4.2.3 UCS Test Results ...... 89
iv 5.5 Conclusions...... 107 5.6 Recommendations...... 109 Chapter 6 Conceptual Design of Preconcentration Waste Handling Systems...... 110 6.1 Introduction...... 110 6.2.1 Rockfill Systems ...... 110 6.2.2 Composite Fill Systems...... 111 6.3 Conceptual Backfill Systems ...... 112 6.3.1 Rockfill...... 112 6.3.2 Composite Fill...... 113 6.3.3 Linking Preconcentration Systems and Backfill Systems...... 114 6.3.3.1 Single Crushing Stage for all of Preconcentration...... 115 6.3.3.2 Crushing Stages for Particle Separation and Reject Disposal...... 116 6.4 Case Studies ...... 117 6.4.1 Thayer Lindsley...... 118 6.4.2 Montcalm ...... 120 6.4.3 Craig...... 122 6.4.4 Fraser Mine ...... 124 6.5 Conclusions and Recommendations...... 127 Chapter 7 Conclusions and Recommendation...... 128 7.1 Conclusions...... 128 7.1.2 Conclusions from Metallurgical Work...... 128 7.1.3 Conclusions from Geotechnical Characterization of Rejects...... 129 7.1.4 Conclusions from Fill Mix Testing...... 129 7.1.5 Conclusions from Conceptual Design of Waste Handling Systems ...... 130 7.2 Recommendations...... 130 References ...... 132 Appendix 1 – Grades for Xstrata Nickel Samples ...... 140 Appendix 2 – Metallurgical Balances...... 142 Craig 8112...... 144 Thayer Lindsley Zone 2 ...... 155 Appendix 3: Assay Values...... 161 Craig 8112...... 162 Craig LGBX ...... 165 Fraser Nickel ...... 168 Fraser Copper ...... 171 Thayer Lindsley Footwall ...... 174 Thayer Lindsley Zone 1 ...... 177 Thayer Lindsley Zone 2 ...... 180 Montcalm East...... 183 Montcalm West ...... 186 Appendix 4: Physical and Geotechnical Properties of Fill Mixes ...... 189
v List of Tables
Table 1. 1: List of Mines and Orebodies Studied...... 4 Table 2. 1: Test plan for each orebody...... 13 Table 3. 1: Summary of Metallurgical Results from Studies and Operations ...... 20 Table 3. 2: Initial Characterization...... 22 Table 3. 3: Particle Size Distribution Summary...... 23 Table 3. 4: DMS Results Summary...... 29 Table 3. 5: Effect of Varying SG Cut on Recoveries for Craig 8112 ...... 31 Table 3. 6: Sorting Results Summary ...... 33 Table 4. 1: Geotechnical Investigation Results...... 46 Table 4. 2: Acid Base Accounting Test...... 54 Table 5. 1: Summary Results from Study of Composite Aggregate Paste ...... 60 Table 5. 2: Results from Composite Fill Studies in Literature...... 61 Table 5. 3: Summary of Physical and Geotechnical Properties for Fill Composed of Pure Rejects ...... 67 Table 5. 4: Summary of Physical and Geotechnical Properties for Composite Fills...... 79 Table 5. 5: Properties of Tailings Samples ...... 81 Table 5. 6: Table of Coefficient of Uniformity, Void Ratio and Specific Gravity for Fraser Copper by Mix Ratio...... 83 Table 5. 7: % Change in Height Between Pouring of Mix and UCS Test...... 85 Table 5. 8: Statistical Analysis of UCS values by Mix Type...... 89 Table 5. 9: Statistical Analysis of UCS / % Cement by Mix Type...... 92 Table 5. 10: UCS and Young’s Modulus at 14 Days (1 cylinder for each test)...... 106 Table 5. 11: Average Values of Key Properties for Each Mix Type ...... 107 Table 6. 1: Preconcentration System Summary for Thayer Lindsley ...... 119 Table 6. 2: Preconcentration System Summary for Montcalm ...... 121 Table 6. 3: Preconcentration System Summary for Craig...... 123 Table 6. 4: Preconcentration System Summary for Fraser Mine ...... 125
vi List of Figures
Figure 1. 1: Map of Xstrata Nickel’s Sudbury Operations (Xstrata Nickel, 2007)...... 3 Figure 2. 1 DMS Vessel...... 8 Figure 2. 2: Apparent Separation Density by Varying Particle Size for Different Slurry Densities at a Flow Rate of 5 lpm ...... 9 Figure 2. 3: Conductivity Sorter...... 10 Figure 2. 4: Showing the slump cylinder and a test specimen with a slump of 75 mm....13 Figure 2. 5: M Test 841 UCS machine ...... 14 Figure 3. 1: Ni Grade and Size Distribution by Size Fraction for Sudbury Contact Ore Deposits...... 24 Figure 3. 2: Cu Grade by Size Fraction for Sudbury Contact Ores ...... 24 Figure 3. 3: Ni Grades by Size Fraction for Sudbury Footwall Ores...... 25 Figure 3. 4: Cu Grade by Size Fraction for Sudbury Footwall Ores...... 25 Figure 3. 5: Ni Grade by Size Fraction for Montcalm Ores...... 26 Figure 3. 6: Cu Grade by Size Fraction for Montcalm Ores...... 26 Figure 3. 7: TL Zone 1 Metal Recovery by Separation Density...... 30 Figure 3. 8: Wash ability Curve for Craig 8112 ...... 31 Figure 4. 1: Size Distribution Curves for DMS Rejects...... 47 Figure 4. 2: Reject Size Distributions vs ASTM Standard Gradations...... 48 Figure 4. 3: Rejects vs Talbot Curves ...... 49 Figure 4. 4: Photograph of 19+13.2mm size fractions for TL Zone 1 (left) and Montcalm West (Right) ...... 50 Figure 4. 5: Void Ratio vs % Flat and Elongated ...... 51 Figure 4. 6: Void Ratio vs 80% Passing Size...... 52 Figure 5. 1: UCS vs Young’s Modulus...... 76 Figure 5. 2: Coefficient of Uniformity vs Young’s Modulus ...... 77 Figure 5. 3: Size Distribution for Tailings and Comparison to ASTM Standard ...... 81 Figure 5. 4: Talbot Curve analysis of maximum density mixes for composite fill...... 82 Figure 5. 5: UCS test cylinders for Fraser Copper...... 84 Figure 5. 6: τττ’ vs Mix Design for Fraser Copper ...... 86 Figure 5. 7: Pictures of Cylinder Slump tests for Fraser Copper ...... 87 Figure 5. 8: UCS vs τττ’...... 88 Figure 5. 9: UCS / % Cement for Composite Fills Made with Cycloned Tailings...... 91 Figure 5. 10: UCS vs % Rejects...... 93 Figure 5. 11: UCS vs % Cement ...... 94 Figure 5. 12: UCS vs. Overall C u of Maximum Density Mixes ...... 95 Figure 5. 13: UCS vs Reject C u...... 96 Figure 5. 14: UCS vs 80% Passing ...... 97 Figure 5. 15: Void Ratio vs 80% Passing Size for Maximum Density Composite Fills ..98 Figure 5. 16: UCS vs Void Ratio ...... 99 Figure 5. 17: Fraser Copper Rockfill at Failure ...... 102 Figure 5. 18: UCS Failure Picture for Fraser Copper Composite Fills...... 103 Figure 5. 19: Young’s Modulus for Full Tailings Composite Mixes...... 104 Figure 5. 20: Young’s Modulus for Cycloned Tailings Composed Mixes...... 104 Figure 5. 21: Young’s Modulus vs UCS...... 105
vii
Figure 6.1: Diagram of Rock Fill System ...... 113 Figure 6. 2: Diagram of Composite Fill System ...... 114 Figure 6. 3: Single Crushing Stage Preconcentration Flow Sheet ...... 116 Figure 6. 4: Independent Crushing Stage for Particle Separation and Reject Disposal Preconcentration Flow Sheet...... 117 Figure 6. 5: Block Diagram of Preconcentration with Mass Flow Based on Run of Mine Ore for Thayer Lindsley...... 120 Figure 6. 6: Block Diagram of Preconcentration with Mass Flow Based on Run of Mine for Montcalm...... 122 Figure 6. 7: Block Diagram of Preconcentration with Mass Flow Based on Run of Mine for Craig ...... 124 Figure 6. 8: Block Diagram of Preconcentration with Mass Flow Based on Run of Mine for Fraser Copper ...... 126
viii Chapter 1 Introduction
1.1 Underground Preconcentration
Underground preconcentration is a concept that involves the use of a proven technology (metallurgical preconcentration) at a unique location in the mining sequence (Scoble et al, 2000; Klein et al 2002). Preconcentration is the rejection of waste from run of mine ores at coarse particle sizes prior to further concentration; this can be accomplished by using simple technologies such as hand sorting and screening or by using highly advanced mineral separation or sorting technologies. A wide range of benefits have been identified for preconcentration including (Schena et al, 1990; Salter & Wyatt, 1991; Feasby and Tremblay, 1995; Peters et al, 1999; Cutmore & Eberhardt, 2002; Klein et al 2002): • Increased mining rate without increasing the size of the fine particles processing facilities; • Reduced grinding and fine particle processing costs by removing hard silicious waste rock; • Increased metal production without increasing the size of the flotation plant; • Reduction in the quantity of fine waste that needs to be disposed of in tailings ponds; and • Separation of non reactive waste from reactive wastes for disposal. The concept of underground preconcentration looks to realize additional benefits from preconcentration by intercepting and rejecting waste earlier in the mining sequence (Scoble et al. 2000; Bamber et al. 2006). A few of the most significant benefits are: • Saving in hoisting and materials handling by rejecting waste as soon as possible in the mining cycle; • Allowing the use of more cost effective bulk mining methods without impacting the downstream grinding and processing system since the process can accommodate higher dilution; • Lowering cut off grade and thereby extending the mine life and increasing resource utilization and
1 • Utilization of rejected material to create economic high strength backfills. It is the expected benefits in utilized rejected material as high strength backfills that this thesis focuses on.
1.2 Objectives
The object of this thesis is to look critically at the waste management implications of underground preconcentration, and specifically at the generation of high quality backfills. There are four phases of the study: • A metallurgical test program to determine ore amenability to preconcentration and estimate the amounts of waste generated; • Characterization of the rejects with respect to their geotechnical properties; • Testing of mix designs utilizing rejects and other metallurgical waste streams; and • Design of conceptual systems for the case study illustrating how the underground preconcentration method would be included in the mining sequence. Xstrata Nickel’s Ontario operations provided the ore samples for the test work and will be used for the conceptual system designs. Individuals conducting further research or designing underground preconcentration based mining systems will benefit from the knowledge developed from this research.
1.3 The Mines and Ores of Xstrata Nickel
Xstrata Nickel’s Ontario operations currently consist of four mines (Craig, Fraser, Thayer Lindsley and Montcalm) and two concentrators (Strathcona and Kidd Creek). In the Sudbury Basin the majority of Xstrata Nickel’s current operations are centered in Onaping where the Craig and Fraser mines and Strathcona mill are located. The Thayer Lindsley Mine is located on the southern rim of the basin approximately 80 km by road from the Strathcona mill. All of the ores mined in the Sudbury basin are processed at the Strathcona mill, resulting in significant transportation costs. The fourth mine is the Montcalm mine located near Timmins, Ontario, which ships ore to the Kidd Creek concentrator about 100 km from the mine site. In addition to the four current mines Xstrata Nickel has three major ongoing projects which could benefit greatly from this work: the Nickel Rim, Onaping Depth, and Fraser Morgan.
2
Figure removed for copy right reasons orginal can be found at http://www.xstrata.com/assets/pdf/xta 20071205_seminar_nickel_slides.pdf
Figure 1. 1: Map of Xstrata Nickel’s Sudbury Operations (Xstrata Nickel, 2007)
Nine orebodies were chosen for study allowing for a broad spectrum of ore types encountered by Xstrata. The table below shows the break down of the ores sampled for this thesis.
3 Table 1. 1: List of Mines and Orebodies Studied Mine Orebody Ore Type Economic Metals
8112 Contact Ni, Cu, Co, PGM Craig LGBX Contact Ni, Cu, Co, PGM
Copper Footwall Ni, Cu, Co, PGM Fraser Nickel Contact Ni, Cu, Co, PGM
Zone 1 Contact Ni, Cu, Co, PGM Thayer Zone 2 Contact Ni, Cu, Co, PGM Lindsley Footwall Footwall Ni, Cu, Co, PGM
East Disseminated Ni, Cu, Co Montcalm West Disseminated Ni, Cu, Co The ores tested in this study can be divided into three different ore types, based on their mineralogical properties (Bamber et al, 2006): • Contact ores are nickel ores with mineralization occurring as massive to disseminated sulfides within the host rock. • Footwall ores consist of a narrow vein high grade stringers containing very high copper grades with significant PGM values. • Montcalm ores are described as consisting of finely disseminated nickel and copper with no economic quantities of PGMs. For Sudbury Basin mines, the results of test work are grouped by deposit type to allow comparison and demonstrate applicability to other deposits in the region.
1.4 Methodology
The four basic phases of this thesis are addressed in four separate chapters. Chapter 2 describes the testing procedures for each phase of the study. Chapter 3 describes different preconcentration methods found in the literature and presents the results of preconcentration testing on the nine ore samples. Chapter 4 describes geotechnical properties of rock and aggregate typically used in fills and presents the properties of the rejects generated from preconcentration testing. Chapter 5 summarizes current practices and research in mix designs and presents the results of testing on mixes using the preconcentration rejects, tailings and cement. Chapter 6 utilizes the test results and
4 literature reviews of the prior chapters in combination with a review of current underground fill systems as a basis for the design of conceptual preconcentration and waste management systems for each of the mines. Chapter 7 presents the overall conclusions and recommendations from this research.
5 Chapter 2 Experimental Procedure
2.1 Introduction
There were four basic phases for this thesis: 1. Sample gathering and classification 2. Metallurgical preconcentration testing 3. Geotechnical classification of rejects 4. Backfill mix evaluation Each of these phases had a distinct campaign of work. Metallurgical testing was conducted using standard metallurgical testing procedures with equipment made available for this testing. For the geotechnical testing and backfill mix evaluation the tests conformed as much as possible with established standards and practices with the largest issue being the result of the small sample size available with which to work.
2.2 Sample Collecting and Classification
2.2.1 Sample Collecting
The sample collecting was done with the assistance and guidance of Xstrata Nickel’s exploration and mine geologists. The procedure was based on their collective experience in sampling and grade control. At the chosen sample point a paint line was drawn on the sample point, then six five gallon steel pails were filled by hand and shovel, taking all specimen along the paint line. Use of the paint line was to avoid sampling bias during collection. The main drawback was that the sample was limited to the surface of the sampling site and as a result limited in the top size of the material that would fit in the buckets; while the sample assay was representative, the size distribution was missing the finest and coarsest fractions of the orebodies.
2.2.1 Mineralogical and Physical Characterization
The samples were characterized on the basis of their observed geological and mineralogical properties. Each of the samples was weighed and its density determined using the displacement method. Hand specimens representing both waste rock and mineralized rock were collected from each sample and kept for future reference.
6
2.3 Metallurgical Preconcentration Testing
2.3.1 Size Assay
Upon completion of the initial characterization each sample was screened into 12 size fractions based on a √2 series with a top size of 254 mm. Each size fraction was then weighed, photographed, and split into two halves. One half was prepared for assay and the other half was saved for subsequent preconcentration testing.
2.3.2 Feed Preparation
Samples were crushed into smaller size fractions prior to pre concentration testing. Due to equipment limitations it was decided to use a top size of 75mm for testing. A jaw crusher with a closed side setting of 64mm was used to crush material larger than 75mm. This effectively split the samples into two portions which were termed as ‘crushed’ and ‘uncrushed’ for the remainder of the testing. Both portions were separated into size fractions of 6.7mm, 25mm+6.7mm, and 75mm+25mm. The 6.7mm size fraction was considered to be too fine for pre concentration and was therefore weighed and assayed. The size fractions were selected to represent the size range that would be processed in an industrial DMS operation and to determine if separation was affected by particle size range. Each size fraction was split into two representative portions so that one portion could be used for DMS testing and the other for conductivity measurements.
2.3.3 Dense Media Separation
DMS tests were conducted using a dense media vessel with ferrosilicon media pumped through the vessel in closed circuit (Figure 2.1). The media density was adjusted by adding FeSi and/or water and the density was checked using a Marcy scale. The ore sample was placed in the vessel and allowed to float or sink. Each sample was separated into four density fractions in order to show the effect of media density on mass rejection and metal recovery. The density fractions were 2.8, 2.95, and 3.1 or 2.7, 2.9, and 3.1 depending on the orebody. The test products were collected, washed, dried, weighed, and assayed. During the washing process it was attempted to recover as much of the FeSi as possible for reuse.
7
Overflow Separation chamber and catch basin for sinks
Catch basin for floats
Circulating pump with adjustable speed
Figure 2. 1 DMS Vessel. (Gray lines show flow of heavy media)
The largest operating constraint for this particular vessel related to the flow rate of the slurry required to maintain a stable suspension. The result of this was that the apparent separation density realized varied in relation to particle size. The apparent density curves for the test were generated based on Stokes Law.
8 20 3.1 2.95 18 2.8
16
14
12
10
8
6
Apparent Separation Densities Separation Apparent 4
2
0 0 50 100 150 200 250 Particle Size (mm)
Figure 2. 2: Apparent Separation Density by Varying Particle Size for Different Slurry Densities at a Flow Rate of 5 lpm
In the graph above there is an increase in the effective separation density for particles smaller than 20 mm. Once particle size becomes less than 10mm the separation density increases at such a rate that no separation occurs. As a result this particular testing unit is only effective from a size range of about 100 mm to 10 mm.
2.3.4 Conductivity Sorting
The conductivity sorting unit used for this testing is on loan from an industrial partner of the University of British Columbia. The rig consists of a conveyor belt which for this test was fed by hand, a conductivity sensor, a processing system, and a pneumatic sorting paddle. The processing system allowed for the sorting criteria used to differentiate between mineralization and gangue to be adjusted slightly for this testing; the machine was adjusted so that nonconductive magnetic particles could be rejected. The sorter was adjusted for the processing of +37.5 mm nickel bearing ores. The design settings of the sorter need to be considered when looking at the sorter results presented in this thesis, due to the smaller particle size and copper content of some of the Xstrata ores.
9
Figure 2. 3: Conductivity Sorter
2.4 Geotechnical Classification of Rejects
Rejects were characterized geotechnically in preparation for their use as fill. The properties considered to be of geotechnical importance were: • Particle size distribution • Particle shape • Adsorption • Specific gravity • Void space • Hardness (strength) • Chemical composition
2.4.1 Particle Size Distribution
Two methods of determining particle size distribution were utilized. ASTM Standard C 136, which provides sieve analysis of fine and coarse aggregates, was utilized for the
10 rejects. A Malvern, a laser particle size analyzer, was used for determining the size distribution of the cycloned and full tailings.
2.4.2 Particle Shape
The main characteristic assessed concerning particle shape was the platy nature of the material under examination. ASTM Standard D 4791, which was followed, provides a test for flat and elongated particles. To ensure a large enough population the size classes measured were +19mm and 19+9.5mm as opposed to size classes based on a √2 series screening. Approximately 100 particles from each size class are measured in three axes; if the longest and shortest axis are found to have a ratio of greater than 5:1, the particle is determined to be flat and/or elongated. Once all the particles in a size class have been determined to have been separated, the weight of the flat and/or elongated particles is taken. The weight of the flat and/or elongated particles divided by the total is the percent flat and/or elongated.
2.4.3 Adsorption and Specific Gravity
For the coarse aggregates the adsorption was determined by ASTM Standard C 127 04. For this procedure a 1 kg sample was split from the main sample. The sample was oven dried and weighted. After weighting it was soaked over night to ensure all internal voids were saturated. Specific gravity was then determined by volumetric displacement in water. Samples were subsequently removed from the graduated cylinder and all the 4.75mm material taken out. Remaining material was then dried quickly by hand to remove any surface water. The +4.75mm fraction was then weighted to determine water adsorption. The specific gravity of fine tailings material was determined using the volume displacement method and a vacuum pump to ensure the removal of trapped air in the graduate cylinder.
2.4.4 Void Space
Void space was measured for both coarse rejects and fine tailings after the specific gravities of each had been determined. The procedure followed closely ASTM Standard C 29. A one liter container was filled and packed with as much material was possible and
11 then weighted. The use of the one liter container was hoped to have a similar container wall effect as the cylinder molds used for the backfill mix testing. With one exception this proved true.
2.4.5 Hardness (Strength)
Due to the limited sample size a destructive sample test was not a possible; as such, the Moh’s hardness scale was used as to approximate hardness. The scale is a standard practice in field geology and consists of comparing the hardness to a common standard. For this test work a copper penny (3), glass scratch plate (5.5), and a hardened steel blade (7) were used.
2.4.6 Chemical Composition
Chemical composition consisted of a sulfur assay as part of the metallurgical assaying done for evaluation of the metallurgical preconcentration stage of testing.
2.5 Backfill Mix Evaluation
There were three basic stages of the backfill mix evaluation phase of test work: mix preparation, slump testing, and uniaxial compress strength testing.
2.5.1 Fill Mixing
Mixes were designed on a material volume basis, then the appropriate masses for each component of the fills were determined. Once the components of a given fill mixture were measured out according to the given recipe all fills were mixed in a similar manner. First all the water and fine material (tailings and cement) were mixed together until the mix was a homogenous slurry or “paste”. The reciepe for the paste was the same for all trials depending on the type of tailings. For cycloned tailings the mix consisted of tailings blended with 5% Portland cement by dry weight at 75% solids. The full tailings were blended with 5% Portland cement at 77% solids. A mixer was used for the cycloned tailings and rockfill cylinders. The mixer was not available for the full tailings due to mechanical issues with the mixer; as a result the mixing was done by hand. The homogenous paste was then poured over the rejects in a large pan and a mortar trowel
12 was used to mix the rejects and the paste. Mixing was considered through when a visual inspection determined that the rejects and fines were thoroughly mixed.
Table 2. 1: Test plan for each orebody
Ore Body Backfill Mix 100% Full Tailings Full Tailings Full Tailings Cycloned Cycloned Cycloned Rejects Max 1:3 1:7 Tailings Tailings 1:3 Tailings 1:7 Craig LGBX X X X Craig 8112 X X X TL Zone 1 X X X TL Zone 2 X X X X TL Footwall X X X X Fraser Cu X X X X X X X Fraser Ni X X X X X X X Montcalm East X X X X X Montcalm West X X X X X X X
2.5.2 Slump Testing
For the slump testing a cylinder with a diameter of 100 mm and height of 150 mm was utilized. In the cases where there was not enough material to fill the 150 mm high cylinder a 100 mm high cylinder was used. The cylinder was filled to approximately half full then a ¼ inch steel rod was used to tamp the sample, after which the second half of the sample was poured. Once full the cylinder was pulled upwards in a steady fashion extruding the fill mix. Slump was measured from the top plane of the cylinder to the top of the extruded material. The outside diameter of the extruded material was also measured.
Figure 2. 4: Showing the slump cylinder and a test specimen with a slump of 75 mm
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2.5.3 UCS Cylinders and Testing
The cylinders for the UCS testing were industry standard plastic molds in 2:1 length to diameter ratios with diameters of 100mm or 76mm. The 100mm cylinders were most commonly used. There were some concerns about the size of the cylinders and this is discussed later in Chapter 5. Filling of the cylinders was done in a similar manner as that discussed for filling the slump cylinder. Once filled the cylinders were tightly covered with impermeable plastic and left for 28 days to cure at room temperature. If necessary the cylinders were capped prior to UCS testing with either sulfur or plaster. Prior to testing the cylinders were weighted and measured. UCS testing itself was done with a M-Test 841 UCS machine.
Figure 2. 5: M Test 841 UCS machine
14 Chapter 3 Metallurgical Results
3.1 Introduction
While the major theoretical focus of this thesis was the use of the rejects, the critical technology of the preconcentration based approach to mining is the means of waste rejection. For this test work dense media separation and conductivity sorting were explored as options. Both methods demonstrated that they were suitable to all the ores tested. The overall results varied by ore types with the footwall ores being the highest performing ores from a preconcentration perspective.
3.2 Literature Review
Chapter one addresses some of the potential benefits of preconcentration and its additional potential if done underground. This literature review will examine the components of a preconcentration plant, and the different technologies applicable to preconcentration. A review of notable research and operational experience in the area of preconcentration will conclude the review.
3.2.1 Preconcentration Plants
Preconcentration plants must perform three basic unit operations, feed preparation, particle separation, and handling of final products (Manouchehri, 2003 and Salter and Wyatt, 1991). Feed preparation is a critical step for the preconcentration plant, since all separation technologies work best within a fairly specific size range. Feed preparation generally consists of reducing run of mine material to the appropriate top size and removing material too fine for the separation stage. Often it is required that several different product sizes must be produced from the feed preparation stage. After the run of mine material has been sized appropriately the material reports to the separation stage. The separation process is generally based on one or more physical properties of the ore, such as specific gravity, color, radioactivity, conductivity, or magnetic. Once the separation has occurred the final operation is a system of conveying the different material to either long term storage for the waste or the next stage of processing for the valuable
15 material. The method of discriminating between valuable and barren material is what constitutes the largest difference between preconcentration systems. Currently, dense media separation and automated sorting haven been identified as the two most promising separation methods for preconcentration.
3.2.1.1 Dense Media Separation Dense media separation makes a separation based on a difference in specific gravity between the valuable and barren particles (Schena et al, 1990). There are two types of dense media separators, dynamic and static. Dyanamic separators use a dense media slurry and centrifugal forces to achieve separation common examples being dense media cyclones and dynawhirlpool. Static separators induce a separation based on the force of gravity alone. The size range of material commonly treated by dense media separation varies based on the method with 1 to 30 mm material treated by dynamic means and coarser material up to +300 mm being treated by static methods (McCulloch et al, 1999). Both static and dynamic processes require a fairly extensive process for mixing and recovering the dense media. Ferrosilicon (FeSi) and/or magnetite slurried with water is the most common heavy media, both of which are magnetic allowing a combination of wet screening and magnetic separation to be used for recovery.
Experience with dense media separation has identified several potential benefits (Munro et al, 1982; Schlitt, 1992; Schena et al, 1990; and McCulloch, 1999):
1. Ability to make sharp separations. 2. Able to quickly adapt to changing production situations. 3. Ability to remove products continuously. 4. Ability to treat a broad range of feed sizes. 5. Maintains separation efficiency through startups and shutdowns.
The major short coming for dense media plants identified in the literature were (Fiedler et. al, 1986 and Bamber 2004):
1. High wear issues in the medium recovery circuit. 2. Poor results in finely disseminated ores.
16 3. Due to the media circulation and recovery circuits, space requirements would be a concern underground. 4. As a wet process, it would require a significant amount of water to be introduced to the underground environment.
3.2.1.2 Ore Sorting
Ore sorting itself is not a new concept, with hand sorting being one of the first methods of minerals processing. Feed preparation is more critical for sorters due the importance of surface characteristics and physical size of the particles most sorters need a 3:1 or 2:1 ratio between the largest and smallest particle to be efficient. Once the particles have been properly prepared for sorting they must be presented to the sensor. To operate efficiently the sensor must be able to analysis each single particle. As a result, feed rate and the materials handling methods are the critical components with this most commonly being done by a conveyor belt or chute (Wotruba, 2006). The critical stage of examining the particle and determining whether material is valuable or barren, is done by a combination of sensor and processing unit. Once the decision of has been made as to accept or reject a given particle, a mechanical device is required to physically make the sort. High pressure jets of air or water and mechanical arms or paddles are generally used to make this separation. Of all the components in a sorter, it is the choice of sensor that controls the design of a sorter.
A multitude of different sensors available and the choice is generally driven by the mineralogy of a given ore. Optical sensors are the most common sensor type, which has been very successfully used in the industrial minerals industry (Wotruba, 2006). After reviewing the literature, if is believed that optical sensors can be considered either passive or active. Passive sensors measure the properties such as color, brightness, reflection, transparency, shape, texture, and size of an ore. Optical sensors are not limited to the visual light spectrum. Conversely, active sensors subject the ore to some form of energy that induces a change that can be detected. An example of an active sensor would be to subject particles to UV light and then detecting fluorescence of selected particles. Another active system involves exposing particles to microwave
17 energy and then using an infrared camera to detect the different heat signatures of materials of different thermal conductivity. Optical sorters require the most intense feed preparation stage often including scrubbing to ensure that the surface of the particles are clean.
Magnetic separators are another form of passive sensor that has been applied in the mining industry. Magnetic sorts are most commonly used for removal of trap steel from crusher circuits. As the name implies magnetic sensor separate particles based on magnetism. Magnetic separators are among the simplest sorting methods, requiring minimal feed preparation and little to no data processing (Vatcha, 2000). Magnetics separators are fairly simple in comparison to many automated sorting, system ultimately accomplish the same task.
Conductivity sensors are a form of passive sensor that detects changes in electrical conductivity. Also referred to as metal detectors, conductivity sensors are best suited for detecting elemental metals and have been identified as having significant potential for sulfide minerals. Conductivity sensor has the ability to detect valuable mineralization that may not be evident on the surface of a particle; this depends largely on the size of particle and the frequency of the sensor. Another common passive sensor type is radiometric, which measures the radioactivity of a material. The most relevant sensor types for this work are conductivity, magnetic, and optical, or a combination of the three.
From the experience gathered from the operational experience and research in sorting some of the benefits associated with sorting are (Wotruba, 2006 and Salter and Wyatt, 1991):
1. Amenable to a wider range of ores, since the magnitude of difference in a given property does not need to be as great, as required for other methods of preconcentration. 2. Sorting can be done without the need for additional water. 3. Two or more methods of sensing can be used simultaneously. 4. Individual sensors are compact and often mobile
18 Some of the challenges identified for sorting in the mining industry from the literature were:
1. The feed preparation requirements for a sorting system can be extensive. 2. The limited size range and feed rate of a single sorter can require multiple sorters to achieve the desired tonnage. 3. Difficulty in servicing the sensors and data processing systems of sorters, due to nature of the suppliers and proper training and retention of maintenance personnel.
When considering these challenges it should be noted that sorters have had success in both the recycling and industrial minerals sectors, where similar issues have been effectively dealt with.
3.2.2 Metallurgical Results from Preconcentration
The results from several preconcentration studies for metallic ores have been published. Table 3.1 summarizes the metallurgical results from studies and operations considered most relevant to this thesis.
19 Table 3. 1: Summary of Metallurgical Results from Studies and Operations Published By Published (Fiedler et (Fiedler al, 1986) (Bamber (Bamber et al, 2006) (Bamber et al, 2006) 91% Ni 97% Ni 98% Cu 85% Cu 95.5% Ag 95.9% Pb 93% PGM 90% PGM 96.6 % Zn Recovery Information u 92.2% Cu et al, 1978) (Miller .8% Ni .8% Ni 29% Cu 29% Cu 8.1% Pb 8.6 Zn % 3.02% Ni 3.02% Ni 1.32% Cu 210 g/t 210Ag g/t Grade of Grade 35 g/t PGM 35 g/t 1.1 g/t PGM 1.1 g/t Concentrate Concentrate to Reporting Concentrator* Concentrator* 55% 22% Mass Mass 27.8% Rejected Head Head 3% Zn3% 55% 6Zn % 94.5% Zn 1976) (Li, Grade 14.8 g/t .39% Ni .39% Ni .98% Ni 38% 1.4% Ni 90% (Vatcha et al, 2000) 6.5% Zn 6.1% Pb 2.54% Ni 2.54% Ni 1.11% Cu 160 g/t 160Ag g/t 13.26% Cu .44 g/t PGM .44 g/t NA 1.65% Cu 83% 9.22% Cu 93.8% Cu (Walter, 1999) 800 3500 6000 Rate Feed Feed tons / day tons / day tons / hour Separation Separation Separation Dense Media Dense Media Dense Media Preconcentration (vein) Pb Zn Sulfide with Ag with Ni SulfideNi Liquid Heavy SulfideNi NA Liquid Heavy NA Ni SulfideNi Magnetic Native CuNative Sensing Conductivity NA 1.66% Cu 67.8% 4.77% C Zn Sulfide Zn Cu SulfideCu Ore TypeOre of Method States Mt. Isa, Mt. Canada Canada Canada Australia Orebody Queensland, Location Location of United States United Sudbury, ON, Sudbury, ON, Sudbury, ON, Sudbury, ON, Sudbury, ON, San AZ, Manuel, MI, United States United MI, Houghton County, Houghton Mascot, TN,Mascot, United Study Study Study Study Study Operation Operation Operation /Operation Zinc Zinc East East Mine Mine Creek Mt.Isa Mt.Isa Operation Copper Copper Whistle Contact Footwall Kingston Kingston American American Company Orebody McCreedy McCreedy McCreedy McCreedy
20 3.3 Procedures
Metallurgical testing consisted of a mineralogical and physical characterization, size assay, dense media separation tests, and conductivity sorter tests. A detailed description of the experimental procedures can be found in chapter two.
3.4 Results and Discussion
As discussed in the procedures the testing program was designed as a means of obtaining preliminary data to be used in a scoping level design for a preconcentration system. After the mineralogical and physical characterization, the tests were ran to simulate a preconcentration plant consisting of sizing of run of mine ore, screening of fines, preconcentration, and finally a reconstitution of preconcentrate and fine for a final preconcentrate that would report to traditional flotation process on surface.
3.4.1 Mineralogical and Physical Characterization
Table 1 summarizes the mineralogical observations and physical properties of the nine ore samples. More detailed descriptions are in the Appendix. Each of the ore samples weighed between 150 and 190 kg and the top sizes varied from 180 mm (7 inches) to almost 300 mm (12 inches). The sulphide mineralization varied from disseminated to massive, although in most cases there was a clear distinction between barren waste and mineralized rock. The SG of hand selected barren waste ranged from 2.85 to 2.98, which was lower than the SG of mineralized rock, which ranged from 3.27 to 4.35. With consideration of the particle size, the density difference is sufficient for gravity concentration using the dense media separation technology. Particle size is a key consideration for gravity consideration, since most methods of gravity separation become less effective at smaller particle size. A good example of this is discussed in Chapter 2 section 2.3.3.
21 Table 3. 2: Initial Characterization
des des des Gangue: No visible sulphide visible No Gangue: sulphides visual No Gangue: Gangue: No visible sulphide visible No Gangue: Gangue: No visible sulphide visible No Gangue: Gangue: No visible sulphides visible No Gangue: Mineralization: Massive pure sulphides sulphides pure Massive Mineralization: Mineralization: Massive pure sulphides sulphides pure Massive Mineralization: Mineralization: Fine disseminated sulphides sulphides disseminated Fine Mineralization: Gangue: Minimal visual fine grained sulphides grained fine visual Minimal Gangue: Gangue: Minimal visual fine grained sulphides grained fine visual Minimal Gangue: Gangue: Minimal visual fine grained sulphides grained fine visual Minimal Gangue: Mineralization: Coarse disseminated sulphides sulphides disseminated Coarse Mineralization: Mineralization: Coarse disseminated sulphides sulphides disseminated Coarse Mineralization: Mineralization: Coarse disseminated sulphides sulphides disseminated Coarse Mineralization: Mineralization: Coarse grained massive sulphides sulphides massive grained Coarse Mineralization: Mineralization: Coarse grained massive sulphides sulphides massive grained Coarse Mineralization: Mineralization: Coarse grained disseminated sulphi disseminated grained Coarse Mineralization: Gangue: Very minimal amounts of fine grained sulphi grained fine of amounts minimal Very Gangue: 1 (mm) Grains Grains Grain SizeGrain Description Mineral No Discrete Discrete No No Discrete Discrete No Specific Gravity Specific Waste Waste Mineralization Calculated Head Grades Head Calculated Size Assay Size Sorter and DMS Top Size Top (kg) (mm) (%) Ni (%) Cu (%) Ni (%) Cu Weight Sample Sample Fraser Ni Fraser 157.6 190 0.669 0.279 0.743 0.378 2.97 3.08 <.1 Ore Body Ore Total Fraser CuFraser 167.3 250 0.753 10.262 0.614 10.939 2.98 4.04 TL Zone 2 Zone TL 1 Zone TL 168.1 190 162.9 0.733 180 0.513 0.582 1.349 0.362 0.755 0.685 0.411 2.89 2.94 3.64 3.23 <1 <1 Craig 8112Craig 157.4 190 1.273 0.508 1.113 0.480 2.96 3.34 <.5 TL FootwallTL 179 195 1.578 8.351 1.245 8.136 2.92 4.35 Craig LGBXCraig 182.2 200 1.922 0.434 2.281 0.326 2.88 3.63 >1 Montcalm EastMontcalm 179.4 230 1.118 0.842 1.637 0.610 2.84 3.61 <1 Montcalm West Montcalm 156.7 300 0.829 0.310 0.369 0.169 2.94 3.27 <.
22 3.4.2 Size Assay
A size analysis was conducted on each sample and the results are summarized in Table 3.3.
The size assay test results were grouped according to the ore type (contact, footwall, Montcalm) in order to compare the results for each ore. Figures 3 to 8 illustrate the copper and nickel grades versus particle size for the deposits in each of the aforementioned groups. The assay results indicated that the distribution of precious metals closely follows the distributions of the nickel and copper.
Table 3. 3: Particle Size Distribution Summary Deposit Top Size P80 P20 -6.7mm (wt %) Craig 8112 190 110 40 3.7% Craig LGBX 200 190 70 1.0% Fraser Ni 190 130 65 2.5% Fraser Cu 250 140 25 11.2% TL Footwall 195 130 55 2.4% TL Zone 2 190 150 65 1.8% TL Zone 1 180 130 30 7.0% Montcalm East 230 160 70 0.7% Montcalm West 300 254 80 1.9%
23 4.5 50% Craig 8112 Grade Craig LGBX Grade Fraser Ni Grade TL Zone 2 Grade 4 TL Zone 1 Grade Craig 8112 Size 45% Craig LGBX Size Fraser Ni Size 40% 3.5 TL Zone 2 Size TL Zone 1 Size
35% 3 30% 2.5 25%
2 % Mass 20% Ni Grade (%) 1.5 15%
1 10%
0.5 5%
0 0%
25 254 190.5 + 127 + 76.2 76.2 + 53.9 53.9 + 38.1 37. 26.5 18.8 13.3 + 9.3 9.3 + 6.6 6 + 4 5 + + 26. + +13 0 19 18 127 0. . .3 5 5 8 Size Fraction
Figure 3. 1: Ni Grade and Size Distribution by Size Fraction for Sudbury Contact Ore Deposits
2.5 50% Craig 8112 Grade Craig LGBX Grade Fraser Ni Grade 45% TL Zone 2 Grade TL Zone 1 Grade 2 Craig 8112 Size 40% Craig LGBX Size Fraser Ni Size TL Zone 2 Size 35% TL Zone 1 Size 1.5 30%
25% % Mass 1 20% Cu Cu Grade (%)
15%
0.5 10%
5%
0 0%
25 254 190.5 + 127 + 76.2 76.2 + 53.9 53.9 + 38.1 37. 26.5 18.8 13.3 + 9.3 9.3 + 6.6 6 + 4 5 + + 26. + +13 0 19 18 127 0. . .3 5 5 8 Size Fraction
Figure 3. 2: Cu Grade by Size Fraction for Sudbury Contact Ores
24 2 45% Fraser Cu Grade TL Footwall Grade 1.8 Fraser Cu Size 40% TL Footwall Size 1.6 35%
1.4 30% 1.2 25% 1 20% 0.8 (%) Mass Ni Grade (%) 15% 0.6
10% 0.4
0.2 5%
0 0% 25 190.5 + 127 76.2 53.9 37.5 26.5 13.3 9.3 + 6.6 254 + 190.5 127 + 76.2 18.8 +13.3 6 4 + 0 + + + + + 53 38 26 18 9.3 .9 .1 .5 .8
Size fraction (mm)
Figure 3. 3: Ni Grades by Size Fraction for Sudbury Footwall Ores
25 45% Fraser Cu Grade TL Footwall Grade Fraser Cu Size TL Footwall Size 40%
20 35%
30% 15 25%
20% 10 (%) Mass Cu Cu Grade (%) 15%
10% 5
5%
0 0%
25 254 + 190.5 190. 127 76.2 + 53.9 53. 37.5 26.5 + 18.8 18. 13.3 + 9.3 9.3 + 6.6 6 + 4 9 8 5 + + + 38. + +13 0 76 127 26 . . 2 1 .5 3
Size fraction (mm)
Figure 3. 4: Cu Grade by Size Fraction for Sudbury Footwall Ores
25 2.5 45% Mont. East Grade Mont. West Grade Mont. East Size Mont. West Size 40%
2 35%
30% 1.5 25%
20% Mass %
Ni GradeNi (%) 1 15%
10% 0.5
5%
0 0%
254 254 190.5 127 76.2 + 53.9 53.9 + 38.1 37. 26.5 18.8 13.3 + 9.3 9.3 + 6.6 6 + 0 5 + + + 26. + +13 19 + 76 18 127 0. .2 . .3 5 5 8 Size Fraction (mm)
Figure 3. 5: Ni Grade by Size Fraction for Montcalm Ores
1 45% Mont. East Grade Mont. West Grade
0.9 Mont. East Size Mont. West Size 40%
0.8 35%
0.7 30% 0.6 25% 0.5
20% Mass %
Cu Grade (%) 0.4 15% 0.3
10% 0.2
0.1 5%
0 0%
254 254 190.5 127 76.2 + 53.9 53.9 + 38.1 37. 26.5 18.8 13.3 + 9.3 9.3 + 6.6 6 + 0 5 + + + 26. + +13 19 + 76 18 127 0. .2 . .3 5 5 8 Size Fraction (mm)
Figure 3. 6: Cu Grade by Size Fraction for Montcalm Ores
26 The size assay for the five ores that constituted the Sudbury Contact ores showed only a slight trend towards enriched fines. This is normally associated with the softer sulphides being more susceptible to breakage and attrition than the host rock.
The footwall ores clearly show enriched fines and significantly lower grades in the coarse fractions. This is best illustrated by Fraser Copper where fractions finer than 13 mm had grades exceeding 20% Cu, which may by pass processing and possibly be added directly to the final flotation concentrate. Also, fractions coarser than 127 mm were practically barren with very low Ni and Cu grades which can be scalped off as a waste product.
Although the results are not as clear for TL Footwall, it is believed that ore in TL muck piles would show similar trends. However, our test sample was taken as a chip sample from the production face and is therefore not representative of muck pile material, and specifically actual particles size distributions.
For the Montcalm ores the assay values are similar across all size ranges. Therefore it is not possible to reject the coarse fraction.
Overall the metal distributions for all ores largely follow the mass distributions. This is most clearly illustrated in the metal distributions for each ore body that can be found in the Appendix.
One of the concerns with regards to the sample gathering protocol (and supported by the observed size distribution) is the low top size and apparent lack of fines. When gathering the samples, material much larger than 250 mm could not be collected because it did not fit into the sample containers. The fines fraction was also missing, as the sample was collected from the surface of the muck pile, where a large quantity of the fines material present had already been washed away. A more representative sample, including the larger fractions of the +254 mm material and the fines fraction, would be required in order to quantitatively determine the potential for scalping of high grade or barren material.
27
Due to the aforementioned limitations of the sampling procedure, predictions of final metallurgical balances would be difficult to determine from these results. The amenability testing results are good indicators of how well these ores will respond to a given pre concentration method. When the metallurgical results are combined with a mineralogical characterization for a specific ore body a reasonable plant design can be projected. The tests provide good estimates of overall pre concentration performance. Using the footwall ores as an example; pre concentration amenability testing results along with estimates of the areas of economic mineralization versus host rock should provide good indicators of ore performance and metal recovery. These indicators would be well within the range of acceptability for a pre feasibility level analysis. This will be more difficult for the contact ores where the economic mineralization and gangue are not as easily discriminated as that found in the footwall ores, but is still possible.
3.4.3 Dense Media Separation
3.4.3.1 Overall Preconcentration
All of the ores exhibited high metal recoveries accompanied by significant mass rejection. The Fraser Copper ore yielded the best results in the DMS study, with nickel and copper recoveries in excess of 96% and mass rejection in excess of 53%. The Thayer Lindsay Footwall ore also yielded very good results, with nickel and copper recoveries greater than 97% and a mass rejection of 37%. The Craig LGBX ore showed a significantly lower copper recovery than nickel, indicating that the copper and nickel are not associated mineralogically (Table 3.4). The copper grade of LGBX was also significantly lower than that seen in other deposits. Preconcentration systems will have difficulty discriminating low grade copper mineralization from barren material if copper is finely disseminated or complexly associated with the host rock.
28 Table 3. 4: DMS Results Summary ) % ( y r e v o c e R 5.712.37 97.634.00 97.18 97.630.69 81.55 91.35 86.89 1.03 96.01 67.83 91.503.58 97.95 97.65 42.70 6.20 97.88 17.49 97.734.39 95.40 43.68 95.656.04 92.60 97.56 71.14 97.57 80.21 93.11 95.25 67.42 82.36 ) % ( e d a r G
. c n o C ) % ( s s a M
. c n o C n o i t a r +2.8+2.9+2.9 75+2.9 47+2.9 63 0.82+2.8 74 0.84 0.48 80 1.83 22.01 1.70 10.79 68 0.82 1.11 0.45 0.55 0.24 a +2.95+2.95 86 68 1.26 3.52 0.57 +2.95 0.38 75 2.12 0.82 p e S 5.66 2.37 4.25 1.83 1.49 3.74 6.22 4.86 5.50 ) % ( e d a r 0.86 G d e e F Ni Cu Mg SG Ni Cu Mg Ni Cu Mg 1.122.46 0.51 0.68 0.31 0.41 0.40 1.19 10.48 1.29 6.99 0.691.62 0.39 0.42 0.66 0.19 t i s o p e Fraser Ni Fraser D Fraser Cu Fraser TL Zone 2 TLZone 1 TLZone Craig 8112 Craig TLFootwall Craig LGBX Craig Montcalm East Montcalm Montcalm West Montcalm
29 3.4.3.2 Precious Metal Recovery
For all the dense media tests precious metals (Au, Ag, Pd, Pt) were assayed and their recoveries tracked. For all the ore bodies the recoveries of Ag, Pd, and Pt followed those of Ni and Cu, usually reporting about 5% lower. Au was more variable; however it is felt this is more a result of the low quantities of Au in the ores resulting in a more variable response due to assay detection limits. The figure below for TL Zone 1 shows the recovery of all the metals assayed by separation density, the variance of Au from the trend for all other metals can be observed.
100%
80%
60%
40%
Metal Metal Recoveries Ni Cu
Co Au 20% Ag Pd Pt 0% 2.8 2.85 2.9 2.95 3 3.05 3.1 3.15 Seperation SG
Figure 3. 7: TL Zone 1 Metal Recovery by Separation Density
3.4.3.3 MgO Rejection
The DMS test results also demonstrate magnesium rejection, indicating another highly promising benefit to pre concentration. Once again, the best results were exhibited by the Fraser Copper ore which exhibited nearly 80% magnesium rejection. The Craig 8112 ore showed the lowest magnesium rejection, with only 13% being rejected at the selected separation SG (Figure 3.8). In all the deposits the magnesium showed a significantly different separation curve from nickel and copper. A clear distinction between the
30 recovery curve for magnesium and those for nickel and copper is exhibited, indicating that it is possible to separate out magnesium preferentially with only minor decreases in metal recoveries.
100%
80%
60%
40% MetalRecovery
Ni 20% Cu
Mg 0% 2.8 2.85 2.9 2.95 3 3.05 3.1 3.15 Seperation Density
Figure 3. 8: Wash ability Curve for Craig 8112
The chart below shows the effect of using 3.1 as a separation density, as opposed to 2.95 for the Craig 8112 Ore Body.
Table 3. 5: Effect of Varying SG Cut on Recoveries for Craig 8112 Separtion Density Mass in Conc. Metal Recoveries % Ni (%) Cu (%) Mg (%) 3.1 43 88.1 84 30 2.95 85 97.3 96.4 86.1
A separation density of 2.95 was chosen in order to separate valuable mineralization and gangue, resulting in nickel and copper recoveries in excess of 96%. However, selecting a separation density of 3.1, while lowering nickel and copper recoveries by 10%, also lead to a significant increase in mass and magnesium rejection. Similar results can be expected for all of the ores tested. Additional testing, using separation densities in the
31 3.1+2.95 range, could yield higher metal recoveries, while still achieving significant magnesium rejection.
3.4.4 Conductivity Sorting
The second major preconcentration method examined for this body of work was conductivity sorting. A detailed description of the test rig and testing program can be found in chapter two. The results for the conductivity sorting were good, though generally not as strong as the dense media separation. It must be noted that the testing equipment was optimized for sorting Ni ores with no precious metals at a fairly coarse size fraction (+25.4mm). As a result ores where the mineralization was concentrated in the smaller fraction or composed largely of copper bearing mineralization did not perform as well, this is especially noted in the Fraser Copper ore (Table 3.6).
3.4.4.1 Overall Preconcentration
The sorter results indicated that conductivity sorting was a viable option for these orebodies. Recoveries for Ni and Cu varied from 59 to 75% with mass rejections ranging from 20% to 70%. The Montcalm West deposit showed the lowest metal recoveries with TL Zone 1 also showing poor metal recoveries, these were the two most disseminated orebodies tested. By ore class, the contact orebodies with the exception of TL Zone 1, showed metal recoveries of 90 93% for Ni and 87 97% for Cu with mass rejections of 17 38%. The footwall ores were more varied showing 81 95% Ni recoveries and 75 88% Cu recoveries with 34 59% mass rejection.
32
Table 3. 6: Sorting Results Summary 40.47 68.22 29.93 ) % ( y r e v o c e R 5.162.39 93.493.73 87.40 95.850.68 67.46 86.70 92.731.08 77.07 89.43 81.123.41 70.67 74.89 94.66 15.42 87.88 90.35 37.51 83.84 59.11 5.584.17 63.076.05 93.60 48.43 85.48 59.23 57.50 ) % ( e d a r G
. c n o C ) % ( s s a M
. c n o C 5.542.574.21 721.81 831.90 80 1.503.41 41 2.436.00 0.57 66 0.944.61 0.37 62 1.655.97 0.40 44 1.85 20.92 75 2.03 12.05 30 0.98 0.87 2.06 0.48 0.64 0.63 0.30 ) % ( e d a r 0.87 G d e e F Ni Cu Mg Ni Cu Mg Ni Cu Mg 1.162.10 0.47 0.81 0.35 0.83 0.36 1.29 11.42 1.40 9.08 0.681.66 0.43 0.32 0.56 0.15 t i s o p e Fraser Ni Fraser D Fraser Cu Fraser TL Zone 2 Zone TL 1 Zone TL Craig 8112 Craig TL Footwall TL Craig LGBX Craig Montcalm East Montcalm Montcalm West Montcalm
33
Precious metal recoveries were similar to those found for Ni and Cu, but were generally not as high as those found with dense media separation.
3.4.4.2 MgO Rejection
There was some evidence that MgO was rejected as part of the sorting process. This appeared to be associated with mass rejection except for the foot wall ores, with a significantly higher level of MgO rejection to weight rejection, indicating that the MgO was associated almost entirely within the gangue portion of the ore.
3.5 Conclusions
The test work showed that preconcentration is a metallurgical possibility for all of the ores tested. While the overall performance of the ores tested varied they all showed that through further refinement high metal recoveries with significant mass rejection was possible. The best results were for those ore bodies with a very distinct differentiation between mineralization and gangue, such as found in the footwall type ores. It is also important to note that the barren material rejected by preconcentration can be associated with one of two sources, gangue that is associated with the ore and dilution from the hanging and footwalls. In the case of the footwall ores, the majority of the material that was rejected was a result of dilution, while the material rejected from the contact and Montcalm ores is largely a result of the gangue within the ore itself.
When DMS and conductivity sorting are compared to each other DMS generally had higher mass rejection and metal recovery, especially for copper. Much of this can be attributed to the differing accuracy and ease of calibration between the two test systems. Irregardless of preconcentration method it can be expected that orebodies of similar mineralogy will respond in a similar manner to preconcentration. For the contact orebodies and Montcalm orebodies, both which were composed of sulfides disseminated through out the ore, preconcentration can be expect to reject 20 40% of run of mine material with metal recoveries greater then 90% with 95% and greater being quite
34 reasonable. For footwall ores, where sulfides are massive and associate with one or two large veins of mineralization, preconcentration can achieve metal recoveries greater than 95% easily. Mass rejection on footwall will be related mostly to the amount of dilution accepted as part of the mining method, so mass rejections ranging from 20 – 80% would be expected. For all three ore types it appears that some magnesium rejection could be expected.
3.6 Recommendations
Based on these results, the following recommendations are made for further testing and evaluation of pre concentration: 1. Collect larger samples to ensure representative size distributions from the muck piles. 2. More in depth DMS testing to determine the “best” crush size and separation density for each deposit. 3. Sorter testing to further investigate conductivity and to assess optical sorting and possibly a combination of optical and conductivity sorting.
35 Chapter 4 Classification of Preconcentration Rejects
4.1 Introduction
The objective of this phase of work was to characterize the pre concentration waste products with respect to their use in backfill. In support of this chapter’s objectives of determining the geotechnical properties and the applicability of preconcentration rejects to use as a backfill, a literature review to determine the appropriate criteria for evaluation was conducted. To facilitate this discussion, a review of the two most logical backfill types and example systems for their use are briefly discussed here. Additionally, the engineering specifications for aggregate use in different roles requiring similar performance were investigated and discussed at the end of the literature review.
4.2 Literature Review
For coarse aggregates in mine backfills the properties most often evaluated are (Stone, 2007; O’Hearn, 2006; O’ Toole, 2004): • Size Distribution • Particle Shape • Void Ratio • Adsorption and Porosity • Specific Gravity • Particle Strength • Particle Hardness and Durability • Chemical Composition The methods for evaluating these properties are largely taken from standard civil engineering practices set forth by the American Society for Testing and Materials (ASTM) or similar organization. It should be noted that in the aggregate industry as a whole, the standards for aggregates varies widely, dependant upon local geologic, economic, and climatic conditions. Mine backfilling is a unique role for aggregates which is not directly comparable to any single role of aggregates therefore, using the criteria for Portland cement concretes is a common starting point for evaluating the
36 components of cemented mine fills. Often the role of minefills is closer to that of civil engineering functions such as stone columns, earth reinforcement, and foundations, than structural Portland cement concrete. The criteria the rejects should be compared to will vary, depending on the type of backfill being used and the method being used at a particular operation. This literature review will conclude with a discussion of which aggregate criteria is considered applicable for preconcentration rejects destined for uncemented rockfill, cemented rockfill, and composite fills.
4.2.1 Particle Size Distribution
The size distribution of aggregates is the property that the engineer has the most control over, usually through crushing, screening, and blending. Void ratio, bulk density, strength, workability, and permeability are among the properties that the particle size distribution directly affects. In civil engineering the size distribution of an aggregate mass is referred to as grading, with well grading being a wide size distribution that results in a dense aggregate mass with minimal voids. Open grading refers to a narrow size distribution with a high void ratio. Another common grading is gap grading, which refers to a size distribution that is wide, but is missing a given size range. The most common method of estimating optimum size distribution is the use of Talbot Curves for civil engineering purposes and mine backfill (Stone, 1993; O’Toole, 2004; Marek, 2001). Originally developed by A. N. Talbot in 1923, Talbot’s curves are used to empirically estimate the size distributions for a mixture of aggregates. The equation is shown below: N Equation 4.1 P= 100 * ( u / u max ) Where: P = percent passing u = opening size
umax = maximum particle size N = distribution constant (Talbot Number) The size distribution of maximum density is generally found in the range of N = .45 to .5. The equation assumes that size distribution is the only property affecting bulk density of a mixture and neglects the role of particle shape and texture, which causes the variance in N values representing maximum density. While Talbot’s curves are used for determining
37 design specifications of size distributions, actual aggregate mixes are often evaluated using the Coefficient of Uniformity. The equation for determining the Coefficient of Uniformity is:
Equation 4.2 Cu = D 60 / D 10 Where:
Cu = Coefficient of Uniformity
D60 = Sieve opening size (mm) through which 60% of aggregate passes
D10 = Sieve opening size (mm) through which 10% of aggregate passes
A value of C u greater than 6 has a wide size distribution (well graded) while a C u of less than 4 has a narrow size distribution (Uniform). When using Talbot’s numbers for size (1/n) distribution C u is equivalent to 6 , meaning a C u of 36 being the equivalent to a Talbot number of 0.5. The specifications for size distribution vary depending upon the end use. The maximum density is often sought when the aggregate mass will be a source of strength, such as road beds and foundations in civil engineering and uncemented and cemented rockfills in mine backfilling. When workability is a consideration, such as Portland cement concretes and mining backfills, a size distribution with excess fines may be desired. When choosing a size distribution, transportation and placement factors must be considered. Mining backfills are often dropped long distances through passes and final placement methods are often less than ideal leading to concerns such as segregation of the aggregate mix and degradation of the individual particles of aggregates. These possibilities need to be considered when evaluating size distributions for aggregates for rockfills in mining (Stone, 2007). In addition to the size distribution, the top size of the aggregate must also be considered. The top size of aggregate used for Portland cement concretes is dictated by workability (White, 2001). For workability of Portland cement concretes, the aggregates must be able to free flow through the form which is being filled; this is rarely be a concern with mine backfills, with the possible exception of underhand cut and fill sill mats. Talbot (1923), White (2001), and Marek (2001) noted that as the top size of aggregate in a given concrete mix increases the economics of the mix. For the vast majority of civil works a top size of 75 to 100 mm is generally the maximum, since
38 compaction, handling, and segregation become problematic for larger sizes. O’ Toole (2004) and Stone (2007), both report segregation problems with aggregate size larger than 75mm. The problem of segregation is commonly controlled by both the size distribution and the shape of particles making up the aggregate mass.
4.2.2 Particle Shape
Like particle size distribution, engineers can exert some control over particle shape, mainly through crushing; however, the level of control is not nearly as exact as for size distribution. This is because the final shape of an aggregate post crushing is dictated largely by the structural geology of the rock mass from which the aggregate originates (Stone, 2007). Aggregate is commonly classified as cubic, blade, disk, or rod shaped in civil engineering practice. Angular cubic particles that are commonly produced by crushing are desired when a high shear strength within an unbound aggregate mass is desired, such as for use in foundations and earth reinforcement (Marek, 2001). The high shear strength of aggregate masses with angular particles is related to the interlocking of particles. The one concern with crushed aggregates is the generation of flat and elongated particles, which can result in segregation and high void ratios all of which have detrimental effects on both the strength and workability of the aggregate mass (Marek, 2001). Flat and elongated particles become a concern when they compose more than 15 25% of the aggregate mass (White, 2001). In the case of Portland cement concretes, the biggest concern with aggregates is their effect on workability. Therefore more rounded aggregates are desirable, since angular particles have poor workability and require more mortar to achieve the same workability. The need for additional mortar increases the consumption of water for a given mix which results in lower compressive strength (White, 2001). Lower compressive strength of Portland cement concretes with angular aggregate is somewhat balanced by increase in flexural strength resulting from the aggregate interlock of angular particles. ASTM Standard D 4791 sets forth a procedure for determining flat and elongated particles based on the aspect ratio (ratio of the largest and smallest dimension of a given aggregate particle in an aggregate mass. A ratio of 5 to 1 is generally considered flat and elongated, though the test is fairly subjective and
39 dependant upon the person conducting the test (Marek, 2001). It is the combination of particle shape and size distribution that control the void ratio in a given aggregate mass, regardless of the presence of a binder.
4.2.3 Void Ratio
The void ratio is one of the single most important properties of an aggregate mass. Minimization of the void ratio results in the maximum density of an aggregate mass; which is the primary goal of the engineer when high strength is desired in an aggregate mass (Marek, 2001). When designing Portland cement concretes, minimizing voids in the aggregate mass often yields the most economic mix (White, 2001). ASTM standard C 29 provides one more accepted means of determining void ratio for aggregate masses. For this procedure the aggregate is placed in a container of a given volume and weighed. If the specific gravity of the aggregate particles is known then the percent voids can be calculated. As long as the properties of the individual aggregate particles are acceptable, variations of the void ratio has the highest impact on the performance of an aggregate mass.
4.2.4 Adsorption and Porosity
Adsorption is the ability of a liquid to penetrate into an aggregate particle (Marek, 2001). Adsorption values for aggregates commonly range from 0 30%, with values of less than 1% being highly desirable. Adsorption is a function of the porosity of an aggregate with a higher adsorption value being an indicator of a more porous aggregate. Generally igneous and metamorphic rocks have a low porosity, while sedimentary aggregates are much more varied (Marek, 2001). For aggregate masses for foundations, porous aggregates commonly have a lower strength and are less elastic. For Portland cement concretes adsorption and porosity have direct effect on the amount of water required in a given mix, which has a direct effect on the water cement ratio (White, 2001.) Stone (2007) and O’ Toole (2004) both wrote about issues of poor strength development related to the moisture content of aggregates adversely affecting the water cement ratio of a
40 cemented rockfill mix. In practice there are several ways to determine the adsorption of an aggregate particle; the most common is by ASTM standard C 127, where the aggregate is allowed to soak for 24 hours in water, after this time the surface of the aggregate is dried and the aggregate weighed. The difference between the weight of the aggregate dry and the wet weight is the adsorption, which is commonly expressed as a percentage of the dry weight. Often the specific gravity provides for quick assessment of the adsorption and porosity.
4.2.5 Specific Gravity
The specific gravity can refer to many different properties of an aggregate, for this thesis it refers to the effective density of an aggregate particle, including internal isolated voids, compared to water. For a given aggregate mass the specific gravity is really only important in determining weights and volumes of aggregate required (Marek, 2001; White, 2001). Specific gravity of aggregates commonly ranges from 2.4 to 3.0, though a higher specific gravity is not necessarily superior to a particle with a lower specific gravity. However, as a general rule an aggregate with a low specific gravity is likely to be more porous (Langer, 1988). Often a high specific gravity is a good indicator that an aggregate will be of high quality with respect to properties such as particle strength and hardness.
4.2.6 Hardness and Durability
Hardness and durability refer to an individual particle’s ability to resist abrasion during handling. Poor hardness and durability can results in degradation of the aggregate mass and cause problems with the handling the aggregates. Soft aggregate particles can break down during handling, changing the size distribution of an aggregate mass and significantly altering the in situ performance (O’Toole, 2004). Additionally, hard aggregates can excessively wear the equipment used to process and handle it. Wear concerns of both aggregate and equipment is especially troublesome in mine backfilling operations where it is common place for aggregate to be dropped hundreds of meters and
41 handled multiple times before final placement. In some operations top size of aggregate for rockfills has been reduced by as much as 50% from the initial processing to final placement (Henderson and Revell, 2005). The most common measure of durability of an aggregate is the Los Angeles Abrasion (LAA) test (Marek, 2001). For minefills it is commonly desired to have an LAA value of less than 20 with a 30 or higher being unacceptable (Stone, 2007). The details of the Los Angeles Abrasion test are described in ASTM Standard C 131. Essentially the test is a partial milling of a carefully graded aggregate mass.
The Moh’s scale of hardness is a common geologic field test for classification of minerals and rocks; it is a suggested field test for aggregates by the Indiana Department of Transportation (INDOT). While INDOT does not specify any particular standards, it is suggested that weakener material can be ruled out as being inappropriate as aggregate.
4.2.7 Particle Strength
Particle strength refers to the stress that an individual particle will withstand prior to failure. The strength of individual particles is of most concern for unbound aggregate masses such as foundations, where high strength of individual particles is desired (Marek, 2001). This is due to the strength of the aggregate mass coming from point to point contact between aggregate particles (Thompson, 2001). In the case of the Portland cement concretes, particle strength needs to be considered for handling and mixing, but is less important for the strength of the final product (White, 2001). Ozturan (1997) stated that for water/ cement ratios greater than 0.4 the aggregate has little effect on normal strength (less the 40 MPa) concretes since the failure is through the cement mortar and cement aggregate bonds and not through the aggregate itself. This is not necessarily the case in cemented minefills where the cement contents are small and point to point contact still plays a major role in the overall strength of the fill. Here individual aggregate particle should have a strength of 70 MPa (Stone, 2007). The strength of particles is often taken from the stress testing of individual particles or taken from the geotechnical tests of drill core in the rock mass from which the aggregate originates (Marek, 2001).
42 The strength of individual particles also needs to be considered when handling the aggregate prior to placement. Along with the individual strength of particles, hardness or durability of the individual aggregate particles is a concern.
4.2.8 Chemical Composition
A chemically stable aggregate is always desired, but not always possible. Chemical composition is more of a concern with cemented fills were the hydration process can result in alkali silicate reactions or sulfate reactions (Marek, 2001). Alkali silicate reactions are seldom a concern with igneous rocks. For this thesis the main concern is sulfide content of the aggregates. For the aggregate industry, sulfides can lead to creation of gypsum in Portland cement concretes, resulting in reduced strength. If sulfides or sulfates are known to be present in an aggregate mix, sulfate resistant cement is normally used, or fly ash is added. For most minefill aggregates it is not a question of whether or not sulfides are present, but rather how much will be present.
4.2.9 How to Compare Minefill Aggregate to Traditional Aggregate
Minefills have a unique role and environment when compared to traditional aggregate. In traditional civil engineering uses aggregates are subjected to many climatic extremes resulting in freeze/ thaw damage, erosion, and temperature related expansion. By and large these extremes are not present in the underground environment. At the same time aggregates used in underground mining are often subjected to extreme geologic conditions such as high stresses and seismic activity. Due to economics, the ability of the engineer to design the aggregate is not nearly as great as in many civil engineering functions. In the case of both cemented and uncemented rockfills the majority of the strength will be provided by the point to point contact of the aggregate itself. The quantity of cement used is usually on the order for 3 5% of total solids; as a result the cements main role is to increase the cohesion of the aggregate mass. The exception is in underhand cut and fill mining where the aggregate mass needs to have sufficient strength to support itself against the force of gravity. With this in mind the requirements for
43 aggregate to be used in rockfills, irregardless of the presence of binders, should be similar to that required for aggregate used for structural support such as road bases and building foundations. Determining which aggregate standard from civil engineering provides the best guide for aggregate utilized in composite fills is more difficult. Composite fills are similar to Portland cement concretes in that a composite fill has a coarse aggregate fraction, fine aggregate fraction, and binder components. The major role of aggregate in Portland cement concretes is as a filler, but there is a significant difference between the water cement and overall cement content for Portland cement concretes and composite fills. The standard that should be used to evaluate aggregate bound to composite fills would largely be a function of the minimum void ratio of the coarse aggregate. When a fill is designed such that the coarse aggregate in the mix will be at or near its minimum void ratio, the coarse aggregate should be compared in the same manner as one would for a rockfill. When the composite fill is mixed so that the void ratio of the aggregate is greater than its minimum, then the same standards as would be used for Portland cement concretes should be used. The theory for this hypothesis is discussed in depth in the literature review of Chapter 5.
4.3 Procedures
The procedures for testing followed ASTM standard tests for the properties examined with some variations. A detailed description of the procedures is can be found in Chapter Two for Geotechnical Classification.
4.4 Results and Discussion
4.4.1 Geotechnical Properties
The rejects from the Dense Media Separation discussed in Chapter 3 were examined to determine the geotechnical properties that would affect their use as a fill material. Due to the small sample size, not all of the data that would normally be obtained was available, so those properties that were felt to be the most critical to a mix design were evaluated. Those properties were: 1. Particle size distribution 2. Particle shape
44 3. Adsorption 4. Specific Gravity 5. Void Space 6. Strength 7. Chemical Composition These properties were selected based on the literature review and consultation with industry professionals. The rejects generally showed results that could be expected from dense igneous rocks subjected to high stress (Table 4.1).
45 Table 4. 1: Geotechnical Investigation Results
d i o V Ratio 8 0.42 .86 0.51 2.78 0.43 29 2.86 0.39 19 2.94 0.44 0.58 2.77 0.41 Adsorption SG
r u f l 43 0.39 2.77 0.40 u S Content % Flat and Elongated and Flat % N u C >7 37.5 22 17 16 8 3.5 4.9 1.1 19.5 16.3 18.3 1.61 0.24 2.95 0.51 5 7 37.5 22 18 16 9.5 5 3.6 1.4 15.0 3.5 10.3 0.72 0.66 2.93 0.49 moh's mm mm mm mm mm mm 19+9.5mm +19mm +9.5mm %Wt. Dry % East West Ore Body Ore Hardness P100 P80 P60 P50 P20 P10 Fraser Ni Fraser 5 7 37.5 20 15 12 4.5 2.5 6.0 1.0 41.2 7.3 27.3 2.07 0. Fraser Cu Fraser 5 7 37.5 22 17 13 5 3 5.7 1.0 26.5 21.6 24.0 0.52 0.29 2 Montcalm Montcalm Montcalm TL Zone 1 Zone TL 2 Zone TL >7 >7 19 37.5 6.5 22 5 15 4.2 15 2 5 1.3 2.5 3.8 6.0 1.3 1.0 82.8 22.1 NA 21.1 82.8 21.7 1.43 0.88 0.82 0.33 2.9 Craig 8112 Craig >7 26.9 8.5 5.4 4.5 1.8 0.8 6.8 0.9 35.3 24.2 34.0 1. TL Footwall TL >7 37.5 23 17 15 5 2.5 6.8 0.9 26.0 24.8 25.4 1.29 0. Craig LGBX Craig 5 7 37.5 15 7 5.5 2.5 1.5 4.7 1.2 12.6 17.4 14.3 1.42
46 4.4.1.1 Particle size distribution
The size distribution is mostly attributed to the processing of the rejects during the dense media trials prior to the commencement of the geotechnical investigation. In practice, the dense media rejects could have a top size that is as high as +75mm. Design of a composite fill that can be transported and placed using a pipeline system dictated a smaller top size. There appears to be two basic envelopes into which the nine orebodies fall: one with a top size equal or greater than 37.5mm that is fairly narrow with few particles below 2mm; and, second, an envelope with a top size of 19 to 37.5mm that is fairly wide at the top with about 20% of particles passing 2mm.
100
90
80
70 Craig LGBX
60 Craig 8112 TL Zone 1 50 TL Zone 2 %Passing 40 TL Footwall 30 Fraser Cu
20 Fraser Ni
10 Mont. East Mont. West 0 0 5 10 15 20 2 30 35 5 Size (mm)
Figure 4. 1: Size Distribution Curves for DMS Rejects
As a comparison, the size distributions for the rejects were plotted against two ASTM C 33 gradation specifications. Gradation number 57 is a gradation of 25mm to 4.75mm and is considered to be a dense grading for civil engineering uses. Gradation number 67 is a gradation from 19mm to 4.75 mm. Both gradations curves would be more appropriate for coarse aggregate used in Portland cement concrete.
47 100 Craig LGBX 90 Craig 8112 TL Zone 1 80 TL Zone 2 70 TL Footwall Fraser Cu 60 Fraser Ni
50 Mont. East Mont. West % Passing % 40 ASTM 67
30 ASTM 57
20
10
0 0 1 10 100
Size (mm)
Figure 4. 2: Reject Size Distributions vs ASTM Standard Gradations
Figure 4.2 shows that six of the nine rejects fall close to the number 57 gradation curve and none fit well into the number 67 gradation curve. The three gradations that do are highly dissimilar to the ASTM gradations and contain a significant amount of finer material. As stated in the literature review the most common way of designing structural aggregate size distributions is by matching to the Talbot’s curve.
48 100 Craig LGBX 90 Craig 8112 TL Zone 1 80 TL Zone 2 TL Footwall 70 Fraser Cu Fraser Ni 60 Mont. East Mont. West 50 Coarse Talbot Curve (N=.5) Fine Talbot Curve (N=.5) % Passing % 40 Coarse Talbot Curve (N=.75) Fine Talbot Curve (N=.75) 30
20
10
0 0 1 10 100
Size (mm)
Figure 4. 3: Rejects vs Talbot Curves
None of the gradations follow a Talbot curve well and it is fairly clear that the rejects are not near a maximum density gradation, shown in the Figure 4.3 by the N = 0.5 curves.
Using the C u for the rejects and the relationship relating C u to Talbot’s Number it was determined that most of the rejects would relate to a Talbot Curve based on an N value of
0.9 to 1.4. The combination of C u and Talbot Number indicates a coarse uniform size distribution for all of the rejects. Craig 8112 and Thayer Lindsley Footwall demonstrated the widest size distributions, while Montcalm West had the narrowest size distribution. By civil engineering standards the reject size distribution displayed here would be just on the verge of being considered dense gradations. To be used as rockfill, adding additional fine material to the mix might be considered to provide a more dense gradation and improve placement characteristics of these rejects.
4.4.1.2 Particle shape
The results of the flat and elongation testing supported observations about the tendency of these ores to fracture in a platy manner. Even after the material had been crushed
49 multiple times, none of the ore bodies met typical concrete standards for particle shape. Fraser Ni and TL Zone 1 material appeared to be the least compliant, but all of the samples had a large population of particles that were just under the criteria for being considered flat and elongated.
Figure 4. 4: Photograph of 19+13.2mm size fractions for TL Zone 1 (left) and Montcalm West (Right)
In the photograph on the left above the flat and elongated nature of the Thayer Lindsley Zone 1 rejects is prominently displayed. Montcalm West (photograph on the right) had the lowest number of flat and elongated particles of all the orebodies. Both orebodies display the angular nature of all the rejects, resulting from crushing, and, at the point the pictures in Figure 4.4 were taken, these particles had been crushed two times. The high level of angularity is highly desirable for structural aggregate uses, since the particles will interlock resulting in a high shear strength. The large number of flat and elongated particles found in several of the orebodies, will be detrimental to the strength properties and void ratios of these rejects. Flat and elongated particles will be prone to breakage and abrasion during handling. Attrition of the flat and elongated particles would need to be further explored, since all of the orebodies’ rejects would benefit from additional fine material in size distributions. Segregation will need to be considered if the rejects are to be placed in a long hole type stope, since there is a fairly narrow gradation and a high population of flat and elongated particles.
50
4.4.1.3 Void Ratio
The void ratio is function of the size distribution and the particle shape. None of the orebodies’ rejects had a size distribution that would produce a maximum density.
0.55
4.9 5.7 0.50 3.6
0.45 6.0 6.0 3.8 4.7
Void Ratio 0.40 6.8 6.8
6.8 6.8 6.0 0.35 6.0 5.7 4.9
4.7 3.8 3.6 0.30 0.0 10.0 20.0 30.0 40.0 50.0 60.0 70.0 80.0 90.0 % Flat and Elongated
Figure 4. 5: Void Ratio vs % Flat and Elongated
The figure above examines the effect particle shape in the form of flat and elongate particles had on overall void ratio. The orebodies were labeled by their C u in the graph to allow orebodies of similar size distributions to be compared. Among orebodies with similar size distributions, an increase in the amount of flat and elongated particles resulted in a slightly higher void ratio. With the exception of the rejects from Montcalm
West (C u 3.6) and Thayer Lindsley Zone 1 (C u 3.8), the reason is believed to be related primarily to the difference in particle sizes involved, since Montcalm West has the coarse overall size distribution of all the rejects and Thayer Lindsley Zone 1 has the finest. Overall, the size of the particles for a given orebodies’ size distribution played a minor role in determining the void ratio.
51 0.55
0.50
0.45
Void Ratio 0.40
6.8 6.8 6.0 0.35 6.0 5.7 4.9
4.7 3.8 3.6 0.30 0 5 10 15 20 25 80% Passing Size
Figure 4. 6: Void Ratio vs 80% Passing Size
A slight bias is seen where a smaller 80% passing size results in a smaller void ratio, however it does not have the same influence as C u or particle shape. For rockfills the highest density possible is desired, which for these rejects would involve some manipulation of the size distribution. Altering the size distributions may require rethinking of upstream processing or the addition of a greater amount of fine material. In the case of composite fills, these aggregates would work well with little alteration being required, since the fine tailings component would provide the material for filling the voids.
4.4.1.4 Adsorption and Porosity
The rejects had low numbers for adsorption, as is usually seen with dense igneous rocks. Low adsorption numbers indicate that any internal pores are poorly connected and would not be expected to adsorb liquids. The low values here indicate that the rejects will have only a minimal affect on the overall water content of any fill mix, since the rejects will not adsorb water needed for transport or binder hydration. This number is for the
52 adsorption of the +4.75 mm material. At the time of testing it was believed that the 4.75 mm material would not pose that much of a concern, since it did not constitute a large amount of the overall mass of the rejects; however, as discussed in Chapter Five, this assumption may have been premature and should be examined further in any future studies.
4.4.1.5 Specific Gravity
Specific gravities ranged from 2.77 to 2.98, which are on the high end of typical aggregates used in civil engineering practices where the SG range is from 2.4 3.0. The high SG of the rejects indicates that they can be expected to be strong and competent with very little in the way of void space within the rejects themselves. This is also supported by the low adsorption numbers.
4.4.1.6 Hardness and Durability
Due to the small mass available for testing, the Moh’s Hardness test was used as an index test to evaluate the hardness and durability of the rejects. The rejects were all fairly hard ranking a 5 or higher on the Moh’s scale. On a subjective level there was very little sign of attrition during the handling of any of the orebodies’ rejects during the course of the test work, which included riffling, screening, and many other activities with the potential to cause significant attrition of aggregates with poor durability. The high hardness does, however, indicate that the rejects will most likely result in fairly high wear rates of the equipment used to handle them if care is not taken in the overall design and operation of any fill system incorporating them.
4.4.1.7 Strength
The strength of the particles themselves was not measured, due in part to the sample size and also the lack of an appropriate sample from which to gather this information. It is believed the rejects from all orebodies were of high strength based on observations during the test work and properties shown to relate to strength. The first observation was that the orebodies’ rejects proved highly resistive to breakage from both manual and mechanical efforts during the test work. The high specific gravities, low adsorption
53 values, and the hardness of the rejects also lend support to the high potential strength of the rejects. Sudbury basin is composed of dense igneous and metamorphic rocks, which commonly have high strengths. For rockfills it is has been stated that aggregates of 70 MPa or greater is appropriate (Stone, 2007), which can be expected of all the orebodies’ rejects. Overall the rejects can be expected to be strong enough that the strength of an individual particle will not adversely affect the in situ strength of a rockfill. For composite fills where the rejects will not be in point to point contact, the strength of the rejects will be of minimal concern.
4.4.1.8 Chemical Composition
The sulfur content of the rejects was determined in two separate tests. Sulfur levels determined from the metallurgical assay is found in Table 4.1, and shows a moderate level of sulfur in all of the orebodies’ rejects. The second test was an Acid Base Accounting (ABA) test, which is used to determine the acid generation potential and the results are displayed below.
Table 4. 2: Acid Base Accounting Test Orebody S(tot) pH NP MPA NNP NP / % Paste Kg/MT Kg/MT Kg/MT MPA
Craig LGBX 1.06 8.285 17.79 33.13 15.33 0.54 Craig 8112 1.05 8.590 13.13 32.81 19.68 0.40 TL Zone 1 1.07 8.454 15.58 33.44 17.85 0.47 TL Zone 2 0.70 8.760 22.09 21.84 0.24 1.01 TL Footwall 0.79 8.225 15.58 24.63 9.04 0.63 Fraser Cu 0.44 8.136 11.78 13.84 2.06 0.85 Fraser Ni 1.43 8.679 12.15 44.69 32.54 0.27 Montcalm East 1.07 8.481 33.01 33.44 0.43 0.99
Montcalm West 0.55 8.541 20.37 17.13 3.24 1.19 Fraser Cu had the lowest sulfur content in both the metallurgical assay and the ABA test. Based on the ABA test there is a potential for acid generation among all the orebodies’
54 rejects. For the purposes of underground fills, the largest concern would be the potential issues related to the binder performance.
4.5 Conclusions and Recommendations
There does not seem to be any overall concerns in the use of these rejects for fills based on their geotechnical properties. While they would not be considered the highest quality aggregates, there would be little significant issue for the use of rejects as an aggregate outside of mine fills either. The largest concern for a fill system will be in the areas of transportation; based on particle hardness and shape there will be a high level of abrasion associated with the transport of the rejects. The in situ performance will be impacted largely by the flat and elongated nature of these rejects, which will pose issues with segregation and attrition during placement and higher void ratios all resulting in low overall strengths. The size distributions of the rejects studied here are fairly coarse and are not as broad as commonly used in civil industry practices. Many of these concerns can be addressed through proper control of the rejects’ size distributions and by the potential to add material such as tailings to broaden the size distributions. The final size distribution of the rejects should be taken into consideration during the design of the preconcentration system. Many mines already have significant processing costs for their backfill components. If these costs are considered when designing the feed preparation for a preconcentration plant, a finer crush of the run of mine ore may become more economical than would be supported by the metallurgical process of preconcentration alone.
55 Chapter 5 Testing of Mix Designs
5.1 Introduction
In chapter four it was stated that the two most logical choices for disposal of preconcentration rejects was as a rockfill or as the coarse component of a composite fill. The two key objectives of this phase of research are to determine which characteristics of the rejects are the most important for mix design and the quality of fill that can be produced with the rejects. Due to limits imposed by the amount of rejects available, this body of work is best viewed as an explorative work, with the objective of helping identify the most critical relationships.
5.2 Literature Review
The literature review begins, by briefly examining current backfill types and performances, with a more focused discussion of theory for design and evaluation of composite fills. The literature review in chapter five looks at the design and evaluation of the backfills, and does not consider the role the backfill distribution system will have on the final performance of the backfills. In practice, once a backfill distribution system is chosen the mix design will most likely need to be adjusted. The role of the backfill delivery systems, and much of the operational experience associated with them, is discussed in Chapter Six.
5.2.1 Current Backfill Practice and Performance
Backfill is a major part of many underground operations for both geotechnical and waste management reasons. There are three basic classifications of fills in current use: paste, hydraulic, and rockfill; the largest means of differentiating between the fills has to do with transportation and composition (Potvin, 2005). Hydraulic fills are reticulated by pipeline under turbulent flow conditions to the mining void with water contents ranging from 60 70% solids by weight (Grice, 2005). Once the hydraulic fill has been placed in the mining void, the excess water must be drained ; this usually results in a void ratio of 1. Most hydraulic fills consist of flotation tailings that have had the finest particles removed. The largest concern with hydraulic fills operationally is high wear rates in pipeline and
56 water management. Hydraulic fills can be delivered as either cemented or noncemented dependant upon the needs of the mine plan. Paste fills are most often pipeline reticulated, though there are a few exception (Brackebusch, 1994). What differentiates paste fills from hydraulic fills is the flow regime and size distribution. Paste fills have a finer size distribution, with a common rule of thumb being 15% passing 20 microns. Water contents for paste fill are low usually 10 25 % by mass (75 90% solids by mass); as a result almost no water bleeds from the fill once placed in the mining void. Due to the fine size distribution and low water content, paste fills are nonsegregating and flow as a laminar fluid during reticulation. Rockfills are a fairly broad category and include complicated highly engineered fills and simplistic unengineered fills (Kuganthan, 2005). Regardless of the sophistication of the system, all rockfills include coarse material, usually waste rock or quarried rock, and are transported to the void by mobile equipment or conveyor. If water is added to rockfills it is generally only for use in cement hydration, though in some cases it has been added for workability reasons. For all three types of backfill, binders are often added to provide increased cohesion and overall stability of the mix, with Portland cement being the most common (Henderson and Revell, 2005). The strength of cement is governed predominantly by the water cement ratio with lower water cement ratios yielding higher strength as long as enough water is present for the full hydration of the cement. Concrete strength is largely a function of the water cement ratio, since enough cement and water is present in the mix to suspend all of the aggregate in the mix, usually 10 – 20% cement by weight (White, 2001). This is not the case in most mine fills where the water cement ratio is higher and the cement content is generally not enough to suspend all the aggregate particles in the mix. As a result mine fills have a lower UCS value than concretes of similar aggregate contents. When examining the strength and binder contents for fills it is important to realize that how binder content is calculated can vary depending upon the operation. Two of the most common ways of calculated binder are either by % of solids or as a % of total mix. The percentage of solids does not consider water when calculating binder content where the % of total mix does. The end result of this is that the strength per given % binder between different operations is not always easily compared.
57 De Souza, DeGagne, and Archibald (2001) conducted a survey of Canadian mines’ fill use in 2001. In this survey the range of binder contents for all fill types ranged from 0.4% to 5.7% cement with an average of 2.9% cement by weight. Part of the survey sought to determine the operational performance of the varying fills by examining the relationship of binder content to UCS strength. The strengths for the backfills ranged depending upon the type of backfill and the amount of binder. The binder content for paste fills ranged from 2 to 5% with a strength range of 0.1 to 1 MPa. Hydraulic fill had binder contents ranging from 2% to 9% with a strength range of 0.5 to 1 MPa. Rockfills had the widest range with a binder content from 3 to 6% and a strength value from 1.5 to 6 MPa. The desire to develop a fill with high strength, high resilience, low void ratio, and tight filling ability has led to research into composite fills.
5.2.2 Composite Fills
To date most work done examining composite fills has been included with rockfills. Common names for composite fills in the literature are cemented aggregate fills, rocky paste fills, and composite aggregate fills (Annor et. al, 2003). Regardless of the name, all composite fills consist of the blending of two or more materials from different sources; and, depending upon the blend ratios, composite fills can have properties closer to paste and hydraulic fills than rockfills. The ability to incorporate both fine tailings and coarse particles makes composite fills a very attractive option for preconcentration based mines (Klein et. al., 2001; Bamber, 2004).
The most in depth study of composite fills has been conducted at McGill University by Alfred Annor, Ferri Hassani and M. Nokken (2003, 2007) in association with CANMET. This work looked at the geotechnical properties of backfills using various blends of mining wastes, in various sample sizes. The study determined that the properties of composite fills are a combination of the tailings fills and cemented rock fill properties. The results discussed by Nokken, Hassani, and Annor (2007) were the most pertinent to this particular thesis. A comparison of the pure tailings, rockfill, and a composite fill of 70% rockfill 30%
58 tailings was preformed on the basis of porosity, UCS, and Modulus of Elasticity. Results from this study are summarized below and additional results are in Table 5.2.
59 Table 5. 1: Summary Results from Study of Composite Aggregate Paste (Nokken et al, 2007). Specimen Binder 7 Day 14 Day 28 Day Deformation Deformation Deformation UCS Porosity UCS Porosity UCS Porosity Modulus Modulus Modulus % dry mass MPa MPa MPa MPa MPa MPa
Tailings A 5 0.29 31.70 32 0.41 31.40 56
Tailings B 5 0.63 35.50 318 1.19 35.30 158 1.34 36.20 408
Rockfill 5 6.33 22.90 1,170 4.37 27.40 12,640
Composite 5 1.12 19.57 830 1.47 17.76 770 1.37 17.26 710
The result summarized above show, that the composite fill had lower uniaxial compressive strength than a rock fill, but had a lower void ratio. A major benefit for the composite fill in this study was that it was able to produce strengths greater than 1 MPa at 7 days with a 5% binder content and demonstrated strain hardening characteristics. Annor (2003) proposes that the ideal composite aggregate paste would be composed of 60 70 % waste rock and 30 40% tailings. The composite aggregate paste fills had several advantages over fills composed of solely fine tailings or waste rock. The benefits for composite fills were: 1. Larger quantity of mine waste can be returned to the underground environment. 2. Higher strength than tailings based fills. 3. Minimal void space resulting in a high resilience.
It is important to note that in the literature review this is the only study that looked at the combination of aggregates larger than 20mm in a composite backfill.
Additional information about composite fills is most often the result of studies conducted as part of mining project evaluations. Since these operational case studies focused mainly on the transportation and operation of fill systems they are discussed in more detail in Chapter 6, however the strength results are summarized in the table below.
60 Table 5. 2: Results from Composite Fill Studies in Literature Operation / Binder Type Binder Content 28 Day UCS Source Components Study and Ratio (% Solids) (Mpa)
Crushed Mullock, Dune Portland Cement and Olympic Dam Baldwin, 2000 1.5 8 0.5 5 sand, Classified tails Fly Ash 1:2
BHP Cannington 30:70 10mm Aggregate Bloss, 2000 Portland Cement 4.8 0.68 (Study) and Dry Tailings
30:70 25.4mm BHP Cannington Bloss, 2000 Aggregate and Dry Portland Cement 4.9 0.66 (Study) Tailings Graded Rockfill and Portland Cement and Kuganathan, Xstrata Mount Isa Classified Hydruallic Copper Reverbatory 4.5 1 2001 Fill (3:1) Furnace Slag (1:2) 12.5mm Crushed Mine Bulyanhulu Landriault, 2001 Waste and Full Plant Portland Cement 3 10 0.5 3.5 Tailings (50/50) 102 mm Waste Rock CANMET Annor, 2006 and Classified Mill Portland Cement 6 4.5 Tailings (70/30) 102 mm Waste Rock CANMET Annor, 2006 and Classified Mill Portland Cement 6 2.83 Tailings (60/40)
The mines that adopted composite fills often did so as a means of developing high strength fills. The addition of the aggregate provided for increased strength by either point to point aggregate contact or by allowing the more efficient use of binders. The more efficient use of binders is accomplished by decreasing the internal surface areas of the fills. To date the design of composite fills has been through empirical trial and error efforts.
5.2.3 Design of Composite fills
Most of the design work in the literature for composite fills has been very empirical in nature. The overall goal is to determine mix designs for composite fills that are able to be transported and placed via pipeline. For this study a more scientific approach to designing composite fills was used. As a first step the current design practices for pumpable concrete were reviewed, due to the similar nature of concrete and pumpable composite mine fills. Much of the literature for pumpable concrete mix design is based on empirical relationships and industry experience and is often dependant on the type of pump used (Kaplan et. al., 2005). These empirical relationships do provide insight into similar problems when considering composite fill mix design. The American Concrete
61 Institute published a technical paper (ACI Committee 304, 1995) that provides a basic empirical design method for pumpable concrete. In addition to the shape and size distribution of the coarse and fine particles, the empirical relationships are based on the top size coarse aggregate, fineness modulus of fine aggregate, binder content, and water content. The goal of the relationship is to ensure that the coarse aggregates are sufficiently coated and suspended within the fine aggregate, cement, and water mixture (often referred to as “mortar”), to facilitate pumping. In practice the relationship restricts the top size of coarse aggregate to a quarter of the pumping systems minimum diameter. The ratio of fine aggregate to coarse aggregate is related to the surface area of the coarse aggregate, where large rounded aggregate requires the least amount of fines and small angular aggregate requires the most fines. Kaplan et. al. (2005) state that a pumpable concrete must remain homogenous (nonsegregating) throughout pumping and must be deformable so as to negotiate elbows and changes in pipe diameter. Kaplan et. al suggested that the bleeding of water from a concrete mix is the largest indicator of poor pumping performance and developed a basic tested, based on a modified air meter, to determine pumpability.
The empirical relationships for pumpable concrete provide significant insight into the design of composite fills, especially with respect to coarse particles. However, there are a few areas of significant difference between composite mine fills and pumpable concrete, namely: 1. The amount of binder in a typical mine fill is 3% dry weight with a water cement ratio of 3:1; where as pumpable concretes often approach 20% cement by dry weight with a water cement ratio of 0.5 to 0.8 2. The size distributions of fine aggregate for backfills are driven by the liberation sized required for metallurgical recovery, resulting in a fine aggregate that is significantly finer than used in the concrete industry. 3. Mining waste products often have fairly narrow size distributions and very angular particles.
62 With the differences between pumpable concretes and composite mine fills in mind , a search was made to determine if a workable model existed on which to start looking at the blending of preconcentration rejects and flotation tailings.
The simplest model for blending of preconcentration rejects and flotation tailings is the work of Ben Wickland et. al. (2006). The model was developed for the co disposal of coarse waste rock and metallurgical tailings. For experimental purposes, blasted and scalped open pit mine waste rock and carbon in pulp cyanide leach tailings were used for Wickland’s work. Based on a literature review of particle packing theory of binary particles, in particular the work of Furnas (1928), the model was developed based on the following assumptions: 1. The mixture is composed of waste rock, tailings, water, and air; 2. Waste rock, tailings, and water are incompressible; 3. The tailings slurry is composed of solid particles and water; 4. The final mixture is homogenous; 5. Mixtures with air may not be homogenous; 6. The mass of air is ignored;
7. There is a large difference in D 50 of waste rock and tailings.
The basic equations for the model are:
Eq 5.1 e = v v / v s
Eq 5.2 er = ( v a + v w + v t ) / v r
Eq 5.3 et = ( v a + v w ) / v t Where: e = Global void ratio
er = Waste rock skeleton void ratio
et = Tailings matrix void ratio
vv = Volume of voids
vs = Volume of solids
63 vr = Volume of waste rock
vt = Volume of tailings solids
vw = Volume of water
va = Volume of air Based on this model Wickland deduced that there are five basic packing types: 1. All waste rock and no tailings; 2. Waste rock with void partially filled with tailings; 3. Waste rock with void space completely fill by tailings; 4. Waste rock suspended or floating in tailings; 5. No waste rock and all tailings. For this model the maximum density is achieved when the tailings would just fill the voids in the waste rock at this point the mechanical properties of the mixture would be controlled by the waste rock, while the permeability would be largely a function of the tailings. As part of this work, the rheology of the tailings fraction was examined; however, the rheology of the entire mixture itself was not critically examined.
5.2.4 Rheological Estimation
Since one of the objectives of this thesis is to design a mixture of preconcentration rejects and flotation tailings which can be reticulated by pipeline, the rheology, or flowability, of the overall mix is a significant consideration. Significant factors affecting rheology for a mine fill are largely the same that affect strength, leading to a strong correlation between strength and rheology (O’Hearn, et. al., 2001). Water content and overall size distribution are the two most important factors, with particle shape playing a lesser role. While there are significant differences in composition for composite fills and pumpable concretes, as discussed earlier, both flow in a similar manner and should meet the same basic criteria of nonsegregation and deformation. In the case of both pumpable concretes and composite fills the mixture itself is carried through the pipe on a thin layer of fine material between the pipe and the mixture being transported. In the case of pumpable concretes this layer is composed mainly of cement and water. For composite fills and mine fills this layer is mainly composed of the very fine tailings fraction (Brackebusch,
64 1994). Currently the ASTM Slump test is the most common method used in industry for measuring fill flowability and rheology (Hallbom, 2005). This approach is likely flawed since it does not take into account all the various factors that effect slump between different mix designs (O’Hearn et. al., 2001). Often the only method of completely ensuring success of a particular mix design is a loop pumping test (Kaplan et. al, 2005). Since a loop test was not feasible for this thesis, another method of testing was required to facilitate a discussion of the potential for a given mix design to be transported via pipeline and to identify potential issues related to the mixes workability.
Based on the sample size that would be available for testing a cylinder slump test was chosen. Cylinder slump tests were first introduced by Chandler (1986), and since then several researchers have developed models to relate cylindrical slump tests to yield strength, most notably Pashias and Boger (1996); Clayton, Grice, and Boger (2002); and Hallbom (2005). Hallbom’s “Lump Model” was chosen for the analysis of cylindrical slump tests in this thesis. The model is a basic numerical model to estimate the shear yield stress of paste and thickened tailings. The model is based on the equation below:
Eq 5.4 τ' = L’ / K e nS’ where: τ' = Dimensionless yield stress L’ = Dimensionless lump K = Constant for failure criteria n = Power constant S’ = Dimensionless slump Two different failure criteria where discussed : the Tresca (K =1/2 n = √3) and Von Mises (K = .577 n = 2). At the time the paper was written the Tresca criteria was used, work since then has indicated that the Von Mises produces better results (Hallbom, 2007).
5.3 Procedures
The primary objective was to explore the utilization of preconcentration rejects for mine with and without flotation tailings. The experimental results presented in this chapter
65 were determined by testing various mix designs using dense media separation rejects and flotation tailings. The mixes were chosen to determine the effects of varying ratios of rejects and cemented tailings mixtures. The composite fills where treated as a binary mixture of rejects with a matrix composed of flotation tailings, cement, and water. The role of the cement content was examined as a secondary consideration for composite fills. Overall the physical, strength, and basic rheological properties of the mixes were examined and compared. Detailed description of the experimental procedure can be found in Chapter 2.
5.4 Results and Discussion
The mix design results for the rockfill trials are discussed separately from those of the composite fills. While some would consider the maximum density mixture of the composite fills to be a rockfill, during the analysis of the data it became clear that the fill mixes containing flotation tailings differed significantly from those mixes containing solely rejects and that they were best looked as two distinct fills. An important note when examining the results is that binder content is calculated on a dry weight basis, where all the water has been removed from the solids. A table that categorically lists all the results for the fill testes by ore body can be found in Appendix 3.
5.4.1 Reject Based Rockfills
A cylinder from each ore body (with the exception of Craig 8112) was tested to determine the amenability of these rejects to use as a reject only rockfill. The mix consisted solely of the dense media rejects and Portland cement. The results indicated that when mixed with binder the rejects produce a rockfill of significant strength. In addition, varying physical traits of each of the orebodies’ rejects were compared to determine what, if any, affect these traits had on strength development.
66
Table 5. 3: Summary of Physical and Geotechnical Properties for Fill Composed of Pure Rejects Orebody Cement Water/Cement Specific Flat and Coefficient of Void Young's UCS Gravity Elongated Uniformity Ratio Modulus
% % +9.5mm Cu e MPa MPa Craig LGBX 5.0 1.00 1.8 14.3 4.67 0.69 235.22 1.47
TL Zone 1 5.0 1.00 1.7 82.8 3.85 0.71 523.42 1.24
TL Zone 2 5.0 1.00 1.7 21.7 6.00 0.74 364.5 1.60
TL Footwall 5.0 1.00 1.8 25.4 6.80 0.62 167.5 2.11
Fraser Cu 5.0 1.00 1.7 24.0 5.67 0.95 264.44 2.11
Fraser Ni 5.0 1.00 1.7 27.3 6.00 0.71 162.8 1.78
Montcalm East 5.0 1.00 1.6 18.3 4.86 1.02 321.93 1.43
Montcalm West 5.0 1.00 1.8 10.3 3.60 0.94 350.52 3.41
5.4.1.1 Physical Characteristics
Overall the physical characteristics of the reject only rockfill samples were similar to those expected in a rockfill. The size distribution had the largest effect on the physical characteristics of the fill composed solely of rejects. The role of size distribution as it pertains to the reject composed rockfills is discussed in chapter 4. When the molds composed of the 100% rejects were examined, the large void ratios and void rations that were expected became evident (see Figure 5.1).
67
Figure 5. 1: Picture of Fraser Copper 100% Reject (rockfill)
It is important to note both the amount of void space and the general size of the voids. As a result of the combination of large voids and high void ratio, the reject only rockfills had a high percolation rate. The percolation rate is relevant to these tests since a considerable amount of the cement slurry was found to have dripped to the bottom of the cylinder, this can be seen in Figure 5.3, despite through mixing of rejects and cement slurry. As such, not all the cement added to the cylinders was involved in the generation of the cylinders final strength. While the water cement ratio was comparable to those found in practice, it appeared to be slightly higher than would have been optimum for these fills.
5.4.1.2 UCS Test Results
Uniaxial compress strength testing showed that the rejects could be used as the aggregate in a rockfill. The overall strengths of the cylinders showed strengths of 1.24 to 3.41 MPa with the average being 1.89 at 28 days. While there was not enough data generated to make any definitive conclusions as to which physical traits had the largest impact on the strength generated by a given orebody’s rejects, there were some traits that appeared to
68 have more influence than others. As expected the size distribution had the most distinguishable effect on the UCS of the all the mixes (Figure 5.2).
4.0
3.5
3.0
2.5
2.0 UCS(MPa) 1.5
1.0
0.5
0.0 2 3 4 5 6 7 8
Reject C u y = 0.2827x + 0.1501 R2 = 0.7027
Figure 5. 2: UCS vs Reject Coefficient of Uniformity
Once the data point for Montcalm West was removed, a rough trend could be established demonstrating an increase in strength with increasing C u. A trend of increasing strength with increasing C u agrees with the literature as discussed in Chapter 4. The increased strength for Montcalm West can partially be attributed to the size of the particles in the distribution. For the rockfills an increase in UCS was associated with a decrease in the % passing 4.75 mm (Figure 5.3) and increase in 80% passing size (Figure 5.4). Unlike the
UCS vs C u trend, the inclusion of Montcalm West does not have a significant affect on the UCS vs % 4.75mm trend or the UCS vs 80% passing size trend. The most plausible explanation for the UCS vs % 4.75mm trend and UCS vs 80% passing size trend is that a
69 larger average particle size results in a decrease in overall surface area. A decrease in surface area would allow the cement to more thoroughly coating the exposed surfaces of the rejects.
4.0
3.5
3.0
2.5
2.0 UCS(MPa) 1.5
1.0
0.5
0.0 0.000 20.000 40.000 60.000
% -4.75mm y = 0.0256x + 2.5095 R2 = 0.3521
Figure 5. 3: UCS vs % 4.75mm
70 4.0
3.5
3.0
2.5
2.0 UCS(MPa) 1.5
1.0
0.5
0.0 0 5 10 15 20 25 80% Passing Size (mm) y = 0.0688x + 0.6305 R2 = 0.3021
Figure 5. 4: UCS vs 80% Passing Size
In addition to the particle size and size distribution, the particle shape appears to also have some affect of overall UCS. A rough relationship between the UCS and the % Flat and Elongated particles can be noticed (Figure 5.5); however, the limited amount of data available here makes it difficult to determine the significance of this relationship. In the literature review of chapter 4 it was stated that until the % flat and elongated particles exceeds 15 – 25% there is minimal effect on the overall strength of an aggregate mass. In the results presented here most the rejects were directly within the range where % flat and elongated particles start to have an effect on the overall strength. With most of the data in this transition zone an overall trend would likely be difficult to determine. It should be noted that the highest UCS (Montcalm West) had the lowest % flat and elongated, while the lowest UCS (Thayer Lindsley Zone 1) had the highest % flat and elongated. With this in mind it appears that results presented here would fit will with
71 4.0
3.5
3.0
2.5
2.0 UCS(MPa) 1.5
1.0
0.5
0.0 0.000 20.000 40.000 60.000 80.000 100.000 % Flat and Elongated y = 0.0142x + 2.2908 R2 = 0.2243
Figure 5. 5: UCS vs % Flat and Elongated
The size distribution, particle size, and particle shape all appeared to have some effect on the overall UCS development in the cylinders examined. These three characteristics have significant bearing on the specific gravity (SG) and void ratio an aggregate mass. In the literature fills with high void ratios and low SGs tended to be brittle with most the strength associated primarily with the cement bonds. Since neither the void ratio (Figure 5.6) nor SG (Figure 5.7) appeared to have a significant effect on the UCS of the reject only cylinders tested, it appears that the effectiveness of the cement bonds control the UCS for the reject only rockfills. The effect size distribution, particle size, and particle shape have on the specific surface area best explains the variation of UCS found in cylinders composed of rejects from different orebodies.
72 3.5
3.0
2.5
UCS(MPa) 2.0
1.5
1.0 0.6 0.7 0.8 0.9 1.0 1.1
Void Ratio y = 0.0759x + 1.7372 2 R = 0.0011 Figure 5. 6: UCS vs Void Ratio
3.5
3.0
2.5
UCS(MPa) 2.0
1.5
1.0 1.60 1.65 1.70 1.75 1.80 1.85 SG y = 1.0406x 0.1108 R2 = 0.0383
Figure 5. 7: UCS vs SG
73 The stress strain curves and photographic evidence also support the theory that the UCS of a reject composed rockfill is primarily a function of the cement bonds. Figure 5.8 shows the controlling failure occurs in the cement bonds in the aggregate mass and not the individual rejects particles. In fact failure in the reject particles was observed in only Thayer Lindsley Zone 2, where an individual particle of mica showed signs of failure.
Figure 5. 8: Fraser Nickel Rock Fill (Left) and Thayer Lindsley Zone 2 (Right) at UCS Testing Failure.
In Figure 5.8 the shear failure plane is clearly evident in the Fraser Nickel and Thayer Lindsley cylinders and there is little other evidence of failure within the cylinder. When the stress strain curves are examined there is a distinct peak that is reached with minimal axial strain followed by a rapid loss of strength and increased axial strain (Figure 5.9). The stress strain curves exhibited the stiff brittle characteristics expected from a fill where the strength is largely a result of the cement bonds.
74 3.5 CR 8112 CR LGBX 3 F Cu F Ni 2.5 MH ML 2 TL Zone 1 TL Zone 2 TL Footwall 1.5 UCS(MPa)
1
0.5
0 0 0.01 0.02 0.03 0.04 0.05 0.06 Axial Strain (mm/mm)
Figure 5. 9: UCS vs Axial Strain curve.
The Young’s Modulus for the rockfill cylinders varied from 162.8 to 523.4 MPa. The modulus varied significantly based on how it was calculated. The values reported in the thesis were calculated by determining the slope of data points from 30% to 70% of the peak UCS value. It is hard to determine a connection between the Young’s Modulus and UCS. A potential trend showing that stiffer fills (Higher Young’s Modulus) had lower UCS values (Figure 5.10).
75 4.0
3.5
3.0
2.5
2.0 UCS(MPa) 1.5
1.0
0.5
0.0 100 200 300 400 500 600
Young's Modulus (MPa) y = 0.0019x + 2.2299 R2 = 0.5037
Figure 5. 1: UCS vs Young’s Modulus
The same factors effecting UCS in an aggregate mass also affect Young’s Modulus, in the literature some relationships between UCS and Young’s Modulus have been discussed, however the results for Young’s Modulus has often been found to highly variable. A trend of decreasing Young’s Modulus with increasing C u is seen in Figure 5.11. Aside from aforementioned trend, at this time not much can be determined about the factors affecting the values of the Young’s Modulus presented here.
76 600.0
500.0
400.0
300.0
200.0 Young's(MPa) Modulus
100.0
0.0 3.0 3.5 4.0 4.5 5.0 5.5 6.0 6.5 7.0
Cu y = 72.687x + 675.28 R2 = 0.4726
Figure 5. 2: Coefficient of Uniformity vs Young’s Modulus
Overall the results for the rockfills indicated that the cement bond plays a significant role in the strength the rockfills, since once the peak UCS was achieved there was a rapid drop in residual strength, indicating that the shear strength of the “reject only fills” was fairly low.
77
5.4.2 Composite Fills
Two types of composite fills were tested: the first was composed of rejects and cycloned tailings, and the second was of full tailings and rejects. The results were similar and are discussed simultaneously. Two design methods were used for the composite fills. The 1:3 and 1:7 mixes are based solely on the ratio of rejects to tailings by volume with the same ratio being used regardless of which orebody’s rejects were utilized. Maximum density mixes were based on Wickland’s method described in Section 5.2.3, as a result the actual mix ratio of rejects to tailings varied depending upon the void ratio of a given orebody’s rejects. The most notable difference between the two was a slightly lower strength with the full tailings composite. This phase of testing focused on the role of reject addition a fixed paste composition. Table 5.4 shows a summary of the results for this test work; it contains the physical characteristics and the testing results for the UCS and Slump tests.
78 Table 5. 4: Summary of Physical and Geotechnical Properties for Composite Fills 247.4 130.6 129.3 208.94 52.464 137.19 49.874 176.77 95.492 124.85 110.02 57.439 76.555 69.832 106.36 97.401 85.709 102.36 75.289 Young's Young's Modulus 13 85 14 14 .19 .67 .66 .85 0.54 0.83 0.24 0.17 UCS / % Cement % 0 0.00 7 0.35 3374 0.54 0.17 75 0.41 54 0.27 .63 0.39 UCS ' τ eMises) (von MPa MPa Void Void Ratio Gravity Specific Water / Water Cement u C of of Uniformity Coefficient Content (%) % % (Dry Mass) (Dry Mix Rejects % Orebody Cement Moisture Fraser Fraser Cu Fraser CuMax CT Fraser Cu 1:3 CT 64.3 Fraser Cu 1:7 CT 36.6 Fraser CuMax Full 2.4 19.8 Fraser Cu 1:3 Full 64.3 4.3 10.7 1:7 Full 5.5 36.6 2.4 17.5 200.21 19.8 20.9 4.3 78.44 4.89 10.7 5.5 6.59 15.9 2.2 4.89 440.85 19.3 247.09 4.84 0.34 4.89 2.1 23.60 0.58 4.37 1.9 0.55 2.4 1.32 4.37 0.67 2.2 0.11 0.29 0.81 2.1 0.07 0.55 0.58 0.70 0 0.64 1. 0.12 0. 0. 0.05 0.7 TL Zone 1TL Zone 1TL Zone Max CT 2TL Zone 1:7 CT 71.5 2TL Zone Max CT 19.4 2TL Zone 1.91:7 CT 69.0Max Full 5.5 19.4 8.7 2.1 76.3 21.0 5.5 70.51 9.4 2.1 6.56 21.0 4.89 217.20 9.4 4.84 2.4 6.49 4.89 69.03 2.1 0.29 2.3 4.84 4.89 0.69 0.58 0.30 2.0 1.66 0.17 2.5 0.58 0.67 0.75 0. 0.28 1.40 0.07 0. 0.58 0 0.79 1.75 0. Craig 8112Craig 8112Craig Max CT 8112Craig 1:3 CT 71.5Max Full 35.9 1.8 71.5 4.0 8.2 1.8 16.3 71.10 8.2 24.71 4.89 129.28 4.90 2.1 4.89 2.0 0.25 2.3 0.45 0.58 0.25 0.26 1.55 0.58 0.98 0 0. TLFootwall TLFootwall Max CT TLFootwall 1:7 CT 76.3Max Full 19.8 1.6 76.3 5.5 7.3 1.6 20.9 238.18 7.3 6.63 4.87 440.13 2.1 4.84 4.87 0.23 2.0 2.2 0.58 0.66 0.23 0.5 0.09 0.58 0.92 0 Craig LGBXCraig LGBXCraig Max CT Max Full 70.7 70.7 2.0 2.0 8.9 8.9 98.54 195.29 4.86 4.86 2.2 2.2 0.28 0.28 0.58 0.58 1.36 0. 0
79 Table 5.4 (Continued) 90.6 91.02 86.68 115.89 64.129 110.01 60.319 113.06 65.065 78.659 109.64 125.15 71.131 84.175 56.875 Young's Young's Modulus 14 .72 .14 UCS / % Cement % 7 0.59 3 0.15 68 0.13 96 0.22 83 0.15 0.91 0.38 1.00 0.39 UCS 6 0.55 0.10 ' 5820 1.16 0.84 0.48 0.20 5807 1.22 0.74 0.48 0.17 τ eMises) (von MPa MPa Void Void Ratio Gravity Specific Water / Water Cement u C of of Uniformity Coefficient Content (%) % % (Dry Mass) (Dry Mix Rejects Fraser Fraser Ni Fraser NiMax CT Fraser Ni 1:3 CT 71.0 Fraser Ni 1:7 CT Fraser Ni 37.1Max Full 2.0 Fraser Ni 20.3 1:3 Full 71.0 4.3 8.8 1:7 Full 5.4 37.2 2.0 17.3 20.3 186.78 20.8 4.3 78.82 8.8 4.84 5.4 5.14 15.8 4.89 2.4 345.15 19.2 248.27 4.84 2.1 0.29 4.84 23.97 4.37 1.9 0.55 0.58 2.4 4.37 0.67 1.44 2.2 0.09 0.29 0.62 2.1 0.07 0.55 0 0.58 0.76 0 0.67 0.11 1.1 0. 0. 0.05 0.8 Orebody Cement Moisture Montcalm EastMontcalm Max CT EastMontcalm EastMontcalm 1:7 CT 63.0Max Full EastMontcalm 20.3 1:3 Full 2.5 63.0 5.4 37.3 11.0 2.5 20.8 4.3 315.89 11.0 4.86 6.57 15.8 720.61 2.3 228.81 4.84 4.86 0.36 4.37 2.0 2.4 0.58 0.67 2.2 0.29 0.11 0.50 0. 0. 0. Montcalm WestMontcalm Max CT WestMontcalm 1:7 CT WestMontcalm 64.8Max Full WestMontcalm 20.2 1:3 Full 2.4 WestMontcalm 64.8 1:7 Full 5.4 37.1 10.5 2.4 20.2 20.9 4.3 326.46 10.5 5.4 4.89 6.56 15.8 709.70 19.2 2.3 232.66 4.84 4.89 0.34 23.63 4.37 2.0 2.4 0.58 4.37 0.67 2.2 0.32 2.2 0.07 0.55 0. 0. 0.67 0. 0.0
80 5.4.2.1 Physical Characteristics
The physical characteristics of the composite cylinders depended on the ratio of rejects to tailings and also on the type of tailing. Both tailings samples were sourced from the Strathcona mill. The full tailings have not been further processed, beyond that required for metallurgical reasons, whereas the cycloned tailings have been cycloned to remove the finest fraction of material. Table 5.5 summaries some of the properties of the two tailings. Table 5. 5: Properties of Tailings Samples SG Bulk Void Fineness Coefficient of Density Ratio Modulus Uniformity
FM Cu Full 2.8 1.55 0.8 0.639 16.6 Cycloned 2.8 1.59 0.8 0.648 6.1 When compared to the fine aggregate used for pumpable concretes these tailings are extremely fine (Figure 5.12).
100 Full Tailings
90 Cycloned Tailings Coarse Limit ASTM C 33 80 Fine Limit ASTM C 33
70
60
50 %Passing 40
30
20
10
0 10 1 0.1 0.01 0.001
Size (mm)
Figure 5. 3: Size Distribution for Tailings and Comparison to ASTM Standard
81 Like the rockfills, the size distribution seemed to have the largest affect on the physical characteristics of the mix designs. When the maximum density mix size distributions are plotted against Talbot Curves of 0.75, 0.50, and 0.25 most of the curves fit within the 0.50 and 0.025 curves (Figure 5.13).
100 Montcalm West Montcalm East 90 Fraser Ni Fraser Cu 80 TL Footwall TL Zone 2 70 Craig LGBX Craig 8112
60
50
% Passing .25 40
30
20 .50
10 .75
0 100 10 1 0.1 0.01 0.001 Size (mm)
Figure 5. 4: Talbot Curve analysis of maximum density mixes for composite fill.
The rejects size distribution was determined by screening and the tailings size distribution by Malvern. Reject and Malvern size distributions were then combined based on the mix ratio mathematically. One of the most notable characteristics of the size distribution curves for all the mixes is the distinct flat portion of the curves found around the 1 mm size range. The flat portion is indicative of the large differences in particle size between the rejects and the tailings. As a result it should be relatively easy to achieve a high density with such mixes. The two curves with the smallest such gap are associated with the Craig orebodies which, for processing reasons, had the finest rejects.
The size distribution had the greatest impact on the efficiency of the particle packing. Packing efficiency is usually determined as a function of the void ratio, where decreasing void ratio indicates more efficient packing. Table 5.6 shows how the C u, void ratio, and
82 specific gravity varied for different mixing ratios. The relationship of the void ratio and specific gravity to size distribution was largely as expected. The widest size distributions had the lowest void ratio and specific gravity.
Table 5. 6: Table of Coefficient of Uniformity, Void Ratio and Specific Gravity for Fraser Copper by Mix Ratio Blend Reject Only Max 1:3 1:7 Tailings Only
Cu 5.67 200.2 78.4 6.59 6.10 Void Ratio 0.95 0.29 0.55 0.64 0.84 SG 1.7 2.4 2.2 2.1 2.0
Unlike the rockfill cylinders, the composite fills showed no sign of segregation. The nonsegregation of the composite fills is attributed to the smaller overall pore size and lower void ratio. Overall the pore size observed in the samples was several orders of magnitude smaller than those observed in the cylinders composed solely of rejects. These smaller pores will result in a very low percolation rate through fills of this nature. Figure 5.14 are photos of composite cylinders composed of Fraser Copper rejects blended with full tailings. The maximum density cylinder is the only cylinder with significant visual difference, were some large voids are noticeable at the surface of the cylinder. These large voids were most likely associated with a container wall effect during the pouring of the cylinder, since these types of voids were not apparent during inspection of the cylinders post failure. The container wall effect unfortunately resulted in some problems with the UCS testing with the Thayer Lindsley Footwall cylinders.
83
Figure 5. 5: UCS test cylinders for Fraser Copper (Top left corner and moving clock wise: full tailings only, 1:7 mix, 1:3 mix, and maximum density mix)
There was evidence of the mixes compressing and bleeding water during the curing process, though it was not until the samples had sat for nearly a day that this behavior was noted.
84 Table 5. 7: % Change in Height Between Pouring of Mix and UCS Test Orebody Rockfills Composites Full Max Full 1:3 Full 1:7 Cycloned Max Cycloned 1:3 Cycloned 1:7
Craig LGBX 2.81 0.00 0.98
Craig 8112 5.26 7.33 1.97
TL Zone 1 1.11 0.00 4.39
TL Zone 2 2.03 7.32 6.83 7.39
TL Footwall 1.20 0.00 9.76 0.00
Fraser Cu NA 5.37 8.29 2.93 3.03 4.39 5.97
Fraser NI 0.94 4.88 4.88 4.15 0.98 3.41 4.88
Montcalm East NA 7.32 7.27 2.56 3.02
Montcalm West NA 4.88 5.12 2.97 2.44 5.37
Average 1.6 3.9 6.8 5.8 2.5 3.3 5.4
Sample of compressibility was observed by the height change. Among the composite fills the maximum density fills had the lowest compressibility indicating that the coarse aggregate was in point to point contact, indicating that the mixes were very near the maximum density possible. The water did not bleed significantly until after the cylinders had been setting for several hours. The time between pouring the cylinders and the observation of bleed water was a few hours, and therefore would meet the requirement for nonsegregation desired for a pumpable mix.
5.4.2.2 Slump Test Results
In combination with the physical observations, the slump test and the resultant rheological analysis allow for general discussion of the potential to pump the composite fill mixes. The “Lump Model” as described in Section 5.2.4 was used to determine a dimensionless yield stress ( τ') which allowed for a comparison of the different mixes. At this stage of testing, there were no apparent trends in the data to indicate which properties governed the values of τ’. The results did show that increasing the amount of rejects increased the τ’. The results did show also that for all of the mixes, segregation was of minimal concern.
85 0.60 Fraser Cu Cycloned Fraser Cu Full 0.50
0.40
0.30 τ'
0.20
0.10
0.00 Max 1:3 1:7 Mix
Figure 5. 6: τ’ vs Mix Design for Fraser Copper
Examining the τ’ for the various ratios of rejects to tailings for both the cycloned and full tailings indicate that for there was no significant difference between the two in respect to τ’. Figure 5.16 shows photographs of the slump tests for mixes of Fraser Copper rejects and cycloned tailings. When Figure 5.16 is examined one can see that that there is a minimal amount of bleed water and segregation apparent in any of the mixes immediately after slumping for any of the mixtures of Fraser Copper rejects and cycloned tailings.
86
Figure 5. 7: Pictures of Cylinder Slump tests for Fraser Copper (Top left clockwise: full tailings only, 1:3 mix, 1:7 ,mix, and maximum density mix)
Any concerns with the small amount of water that later bleed from the mixes would require only a slight adjustment in the water content to alleviate. In the case of the maximum density mix there is very little noticeable slump, which might be grounds for a change in mix design, since such a stiff fill would require high pressures to pump. Unfortunately, there were no discernable physical traits that could be connected to the τ’ based on this series of testing. A potential trend of increasing UCS with increasing τ’ is noticeable, which indicates that future testing, or more extensive testing, may show a relationship (Figure 5.17).
87 2.0 Cycloned Full
1.8
1.6
1.4
1.2
1.0 UCS
0.8
0.6
0.4
0.2
0.0 0.000 0.100 0.200 0.300 0.400 0.500 0.600 τ'
Figure 5. 8: UCS vs τ’
Future work will likely require testing of larger specimens than those tested here in order to exceed the yield stress for some of the mixes tested here.
88
5.4.2.3 UCS Test Results
Uniaxial Compress Stress testing showed several important facets of how the addition of the rejects affect strength development in the composite fills. A brief statistical analysis comparing each overall mix design based on the UCS test results showed high variations in the maximum density cylinders and less variation for the 1:3 and 1:7 mixes (Table 5.8). The maximum density cylinders had the highest average strengths for the composite fills, but were still significantly weaker than the rockfill samples.
Table 5. 8: Statistical Analysis of UCS values by Mix Type Mix Type # of Maximum Minimum Average Standard Cylinders Deviation MPa MPa MPa MPa RockFill 8 1.24 3.41 1.89 0.69 CT Max 9 0.57 1.66 1.24 0.35 CT 1:3 3 0.62 0.98 0.80 0.18 CT 1:7 7 0.68 0.92 0.77 0.08 Full Max 8 0.54 1.75 1.07 0.41 Full 1:3 4 0.74 0.96 0.82 0.10
Full 1:7 3 0.55 0.83 0.70 0.14
Some of the variation for the maximum density cylinders was a result of poor specimens, namely Craig LGBX, Criag 8112, and Thayer Lindsley Footwall for the full tailings and Thayer Lindsley Footwall for cycloned tailings. Craig LGBX and Craig 8112 had problems with the full tailings mix largely as a result of the mixing being done by hand. As a result the tailings, cement, and water were likely not as throughly mixed. The poor results for Thayer Lindsley Footwall are mostly attributed to a reject void ratio that varied significantly from that measured during the geotechnical evaluation discussed in Chapter 4. As a result, the mix ratios for Thayer Lindsley Footwall were deficient in tailings resulting in large voids in the cylinder (Figure 5.18).
89
Figure 5. 18: Thayer Lindsley Footwall Rejects and Full Tailings Maximum Density Cylinder.
Even when the cylinders with procedural problems are considered, the fills where the rejects were in point to point contact (rockfills and maximum density) exhibited higher degree of variability in UCS than the 1:3 and 1:7 mixes. Due to the varying cement contents in the mixes, the UCS per unit of cement (UCS / % cement) of all the mixes was determined as means of evaluating the “efficiency” of the cement in a given mix (Figure 5.19 and Figure 5.20). The maximum density cylinders proved to be the most efficient mixes based on the UCS / % Cement.
90 0.90 RockFill
0.80 Full Max Full 1:3 0.70 Full 1:7
0.60
0.50
0.40 UCS/ Cement % 0.30
0.20
0.10
0.00 Craig Craig TL Zone 1 TL Zone 2 TL Fraser Cu Fraser Ni Montcalm Montcalm Average LGBX 8112 Footwall East West Orebodies
Figure 5. 19: UCS / % Cement for Composite Fills Made with Full Tailings
0.90 RockFill CT Max 0.80 CT 1:3 CT 1:7 0.70
0.60
0.50
0.40 UCS/ UCS/ Cement %
0.30
0.20
0.10
0.00 Craig Craig 8112 TL Zone 1 TL Zone 2 TL Footwall Fraser Cu Fraser Ni Montcalm Montcalm Average LGBX East West Orebodies
Figure 5. 9: UCS / % Cement for Composite Fills Made with Cycloned Tailings
91 A statistical analysis of UCS / % cement by mix type shows a noticeable difference in average UCS / % cement between the maximum density mixes and 1:3 and 1:7 mixes for both full and cycloned tailings (Table 5.9). There is also a difference in the degree of variability between mix ratios. The mixes composed of 1:3 and 1:7 rejects to tailings are not highly variable; however the maximum density mixes exhibit a high degree variability.
Table 5. 9: Statistical Analysis of UCS / % Cement by Mix Type Mix Type # of Maximum Minimum Average Standard Cylinders Deviation MPa MPa MPa MPa
RockFill 8 0.68 0.25 0.38 0.14 CT Max 9 0.85 0.35 0.60 0.20 CT 1:3 3 0.24 0.14 0.19 0.05 CT 1:7 7 0.17 0.13 0.14 0.01 Full Max 8 0.83 0.27 0.50 0.17 Full 1:3 4 0.22 0.17 0.19 0.02
Full 1:7 3 0.15 0.10 0.13 0.03
A noticeable shift in the trend of UCS vs % rejects occurs when the rejects come into point to point contact as opposed to being suspended within the tailings cement paste. Figure 5.21).
92 2.0 Cycloned 1.8 Full
1.6
1.4
1.2
1.0
UCS(MPa) 0.8
0.6
0.4
0.2
0.0 0.0 20.0 40.0 60.0 80.0 100.0
% Rejects
Figure 5. 10: UCS vs % Rejects
It is interesting to note is that the increase in UCS with increasing % rejects also coincides with an increasing UCS with decreasing cement content (Figure 5.22).
93 2.0 Cycloned 1.8 Full
1.6
1.4
1.2
1.0
UCS(MPa) 0.8
0.6
0.4
0.2
0.0 0.0 1.0 2.0 3.0 4.0 5.0 6.0 % Cement
Figure 5. 11: UCS vs % Cement
These trends are supported by the literature if it is accepted that in the maximum density mixes the coarse rejects are carrying the load, whereas in the 1:7 and 1:3 mixes the cemented tailings are carrying the load. Based on this premise, the high variation of UCS within the maximum density samples is a result of the reject properties. Since a composite fill at maximum density would fit most definitions for a rockfill, it would makes sense that factors that proved most influential to the rockfill strengths; size distribution, particle size, and particle shape, would also be true for the maximum density composite fills. The UCS vs Overall C u, however showed a trend of decreasing UCS with increasing C u, unlike the rockfill mixes (Figure 5.23). The trend was fairly significant for the cycloned tailings mixes.
94 2.0 CT Max 1.8 Full Max
1.6
1.4
1.2
1.0
UCS(MPa) 0.8
0.6
0.4
0.2
0.0 0 50 100 150 200 250 300 350 400 450 500 550 600 650 700 Overall C Full Tailings u Cycloned Tailings y = 0.0005x + 1.5728 y = 0.0023x + 1.7535 2 R = 0.1012 R2 = 0.8261
Figure 5. 12: UCS vs. Overall C u of Maximum Density Mixes
The C u of the rejects seemed to have a fairly minor if any role in the overall strength development of maximum density composite fills (Figure 5.24).
95 2.0 CT Max Full Max 1.8
1.6
1.4
1.2
1.0
UCS(MPa) 0.8
0.6
0.4
0.2
0.0 2.5 3.5 4.5 5.5 6.5 7.5
Reject C u
Figure 5. 13: UCS vs Reject C u
While, the size distribution of the rejects doesn’t have seem to have a significant effect on the UCS the particle size of the rejects does seem to have more an influence (Figure 5.25). There is a distinguishable trend for the cycloned tailings that correlates a decrease in 80% passing with an increase in UCS. A similar trend is not as apparent for the full tailings mixes, but it is conceivable that a similar affect could be expected.
96 2.0 Cycloned 1.8 Full
1.6
1.4
1.2
1.0
UCS(MPa) 0.8
0.6
0.4
0.2
0.0 0.0 20.0 40.0 Full Cycloned 80% Passing (mm) y = 0.0412x + 1.8966 y = 0.037x + 2.0493 2 R2 = 0.5518 R = 0.1541
Figure 5. 14: UCS vs 80% Passing
The decreasing UCS with increasing 80% passing size is most likely associated with the container wall effect. General practice dictates that a cylinder size of 3 – 10 times the largest particle is required to avoid the container wall effect. The container wall effect would result in high void ratios due to interactions between container wall and the rejects. An increase in void ratio with increasing 80% passing size is clearly evident in the maximum density composite fills which can be seen as a strong indicator that the container wall effect influenced the UCS results for the maximum density mixes (Figure 5.26).
97 0.38 Cycloned Full 0.36
0.34
0.32
0.30
0.28 Void Ratio Void
0.26
0.24
0.22
0.20 0 5 10 15 20 25 30 Full 80% Passing (mm) Cycloned y = 0.005x + 0.1984 y = 0.0041x + 0.2424 R2 = 0.9632 R2 = 0.5452
Figure 5. 15: Void Ratio vs 80% Passing Size for Maximum Density Composite Fills
The effect of the void ratio on the UCS was most clear for the maximum density mixes composed of rejects and cycloned tailings where an increase in void ratio corresponded with a decrease in UCS (Figure 5.27). A trend of higher UCS with lower void ratio was less clear in maximum density mixes containing full tailings.
98 2.0 Cycloned
1.8 Full
1.6
1.4
1.2
1.0
UCS(MPa) 0.8
0.6
0.4
0.2
0.0 0.20 0.25 0.30 0.35 0.40
Full Tailings Void Ratio Cycloned Tailings y = 10.113x + 4.3131 y = 5.5177x + 3.0147 2 R2 = 0.2945 R = 0.6115
Figure 5. 16: UCS vs Void Ratio
In addition to the void ratio, the specific gravity is often used to determine the efficiency particle packing with a higher specific gravity corresponding to a lower void ratio. When the UCS and specific gravity of the differing mixes are plotted against each other, a clear trend of increasing UCS with increasing SG for the maximum density mixes with full is seen (Figure 5.28).
99 2.0 Cycloned 1.8 Full
1.6
1.4
1.2
1.0
UCS(MPa) 0.8
0.6
0.4
0.2
0.0 1.5 1.7 1.9 2.1 2.3 2.5 2.7 2.9 Full Tailings SG Cycloned Tailings y = 4.772x 10.178 y = 1.5873x 2.3367 2 R = 0.8912 R2 = 0.2168
Figure 5. 28: UCS vs SG
Based on the results presented here the peak UCS for composite fills consisting of mix ratios 1:3 and 1:7 is controlled by the interactions between the tailings and cement. Once the ratio of rejects to tailings approached maximum density the rejects became the controlling property of the composite fills. The packing density was the critical factor controlling peak UCS values. The packing density’s effect on UCS was best exhibited by the UCS vs void ratio for cycloned tailings and UCS vs SG for full tailings. The true peak values of the maximum density mixes tested during for this thesis were likely not achieved due to limitations caused by the size of cylinders used for testing. The particle size was the controlling factor for packing efficiency for the maximum density mixes tested due to the container wall effect, if larger cylinders are tested the size distribution can be expected to be a more dominate factor.
100
As part of the UCS testing, stress strain curves were obtained for the mix designs. In comparison to the rockfills discussed earlier, the peak UCS values were lower for all of the composite fills; however, the composite fills had distinctively different stress strain curves. (Figure 5.29). Composite fills maintained strength post failure over a wider range of strains indicating a higher resilience.
2.5 F Cu Rockfill F Cu Full Max F Cu Full 1:3 2 F Cu Full 1:7 Full Tailings
1.5
UCS (MPa) 1
0.5
0 0 0.01 0.02 0.03 0.04 0.05 0.06 Axial Strain (mm)
Figure 5. 29: UCS vs Axial Displacement Curves for Fraser Copper Rockfill and Full Tailings Composite Fills
The rockfills typically exerted a stiff brittle failure with the fairly clear failure plane being evident through the cylinders. Often the two halves of the sample remained largely intact (Figure 5.30).
101
Figure 5. 17: Fraser Copper Rockfill at Failure
In comparison, the tailings based fills exhibited a yielding failure with the no clear failure planes, often exhibiting bulging and flaking along the sides of the cylinder. The maximum density cylinders exhibited a more yielding failure than the rockfills and maintained strength over a much wider range of strains, indicating that individual components of the fill were providing the strength as opposed to just the cement bonds. than the rockfills in post failure photographs (Figure 5.31).
102
Figure 5. 18: UCS Failure Picture for Fraser Copper Composite Fills (Top left: Full Tailings; top right: 1:7 mix; bottom left: 1:3 mix; bottom right: maximum density)
The Young’s Modulus for the composite fills was calculated based on the stress strain curves as an indicator of the fills relative stiffness. The calculated Young’s Modulus varied and it was difficult to see many trends in the information as a result. Three noteworthy trends based on the results shown in Figures 5.32 through 5.34 were:
1. Overall the cycloned tailings fills were stiffer than a full tailings fill 2. An increasing ratio of rejects to tailings produced a stiffer fill when full tailings were used. 3. A higher UCS value corresponded with a stiffer fill.
103 600 Rockfill Youngs Full Max Youngs 500 Full 1:3 Youngs Full 1:7 Youngs
400
300
200 Young'sModulus (MPa) 100
0
2 ll g 1 MH ML F Cu F Ni Av Zone 1 CR 81 CR LGBX TL TL Zone 2 TL Footwa
Figure 5. 19: Young’s Modulus for Full Tailings Composite Mixes
600 Rockfill Youngs CT Max Youngs CT 1:3 Youngs 500 CT 1:7 Youngs
400
300
200 Young'sModulus (MPa)
100
0
l 2 u H 1 2 l 1 BX M ML e e F C F Ni twa Avg 81 LG Zon Zon CR L L Foo CR T T L T Figure 5. 20: Young’s Modulus for Cycloned Tailings Composed Mixes
104 300 CT Max Full Max
250 CT 1:3 CT 1:7 Full 1:3 Full 1:7
200
150
100 Young's (MPa) Modulus
50
0 0.0 0.5 1.0 1.5 2.0 UCS (MPa)
Figure 5. 21: Young’s Modulus vs UCS
5.4.2.4 Effect of Cement on Composite Fills of Maximum Density Mixes
Montcalm West and Fraser Copper rejects were used investigate the effect cement content had on a composite fill’s properties at maximum density (Table 5.10). Montcalm West and Fraser Copper show conflicting results as to whether additional cement increased peak UCS values; however, both exhibited increased stiffness with increased cement content. An interesting note is that the Fraser Copper mix in this test had a higher UCS value than a 28 day sample of comparable cement content.
105
Table 5. 10: UCS and Young’s Modulus at 14 Days (1 cylinder for each test) Orebody Cement Youngs UCS Content % MPa MPa Montcalm 2.1 64.595 0.685 West Montcalm 3.0 119.85 0.729 West Montcalm 3.9 292.04 1.843 West
Fraser Cu 2.3 140.58 1.718
Fraser Cu 4.1 238.53 1.654
106
5.5 Conclusions
Overall the most significant conclusion is that the incorporation of preconcentration rejects into a mine’s fill scheme is plausible and likely an option that will provide significant advantages over conventional fills. There were some limitations due to the sample size of rejects available for testing that will need to be addressed in any future testing, despite these limitations some meaningful observations were made (Table 5.11).
Table 5. 11: Average Values of Key Properties for Each Mix Type Mix # of UCS / Std. τ' Std. Void Ratio Std. SG Std. Cylinders %Cement Deviation Deviation Deviati Deviatio RockFill 8 0.38 0.14 NA NA 0.80 0.15on 1.73 0.07n CT Max 9 0.60 0.20 0.58 0.00 0.30 0.04 2.26 0.10 CT 1:3 3 0.19 0.05 0.15 0.10 0.51 0.06 2.05 0.01 CT 1:7 7 0.14 0.01 0.09 0.04 0.67 0.01 1.98 0.07 Full Max 8 0.50 0.17 0.58 0.00 0.28 0.03 2.34 0.11 Full 1:3 4 0.19 0.02 0.13 0.05 0.53 0.02 2.20 0.03 Full 1:7 3 0.13 0.03 0.05 0.01 0.66 0.02 2.16 0.07
1. A higher UCS / % Cement ratio indicates a more efficient use of binders, meaning that higher strengths can be achieved while using less cement for fills composed of both rejects and flotation tails as opposed to fills composed of solely rejects or flotation tails.
2. The increase in specific gravity and a decrease in void ratio indicate a denser packing of the material. A denser packing should result in a stronger fill that maintains strength over a wider range deformation.
3. The one possible drawback that was identified in this work was an increase in τ’, which is a measure of workability, meaning that transportation to the stope and tight filling in the stope with a composite fill may be more energy intensive than that of a fill composed solely of flotation tails.
4. When compared to current industry performance for rockfills and paste fills the reject based fill performance seen in this study are comparable. Rockfills currently achieve strengths of 2 to 7 MPa which was higher than the strengths
107 seen in the study, but not significantly so when the cylinder sizes tested are considered. Strong paste fills are usually in the 1 – 2 MPa range, which the composite fills tested here achieved.
108 5.6 Recommendations
This work is at best viewed as an explorative or amenability level study, a more intensive study would be required to mathematically determine the relationships of the differing properties. In this test there was a constant recipe of flotation tails, water, and cement; with only the ratio of this “paste” to rejects varying. This was done to ensure that the key variable studied was the addition of the rejects. Future work should focus on varying the overall mix recipes. There is a growing body of literature in both the areas of mine fills and concrete related to blending materials of different size distributions, meaning much of the science necessary has been developed. The main shortcoming in the literature was in the transport of composite fills most studies and operations to date have limited aggregate addition in pastes to around a top size of 20mm and less than 50% of the total mix by weight. To gain the maximum strength benefits from composite fills the upper limit on the top size and percentage of the total mix would need to be increased; however, for many mines these high strengths may not be required. Most likely the top size would need to approach 37.5 mm and the total composition would need to approach 60 70% coarse rejects by volume. A possible answer to the transportation issues surround composite fills with high aggregate contents may be a fill that is mixed for a ratio of rejects to paste that is slightly less than necessary for point to point contact of the rejects. Such a fill would likely exhibit strain hardening as the initial strain would bring the rejects into the point to point contact that generates the high strengths of composite fills at maximum density. A key requirement for any work along these lines will be a large sample of rejects, preferably several tons of rejects, along with an equal amount of flotation tails. Future work should also consider the possibility of adjusting the size distribution of the individual components of each fill in addition to the mix ratios.
109 Chapter 6 Conceptual Design of Preconcentration Waste Handling Systems
6.1 Introduction
Based on the work presented in the preceding three chapters and a brief review of current backfill systems, this chapter presents preconcentration systems for Xstrata Nickel’s Ontario mines. Current rockfill and composite fill systems designs were reviewed. Based on these designs as well as considerations of the characteristics of the rejects and fill mix test results system for preconcentration and backfill preparation were designed.
6.2.1 Rockfill Systems
Rockfill systems consisted of two basic approaches; one, a highly engineered approach using quarried and graded aggregate and, the other, a less rigorous approach involving the use of development waste. Barrick’s Meikle Mine (Sacrison, 2001) uses quarried and graded rock fill in combination with binder that is mixed in an underground plant and then handled by mobile equipment. A similar system is used at Mt. Isa’s George Fisher Mine; however, the rock is 16mm rejects from a dense media separator (Kuganathan, 2005). For both mines, the material is sent underground through a series of boreholes and passes. In both these systems the aggregate is precisely measured and mixed in a highly automated and controlled process. The sophistication of the Mt. Isa and Meikle systems was associated with underhand cut and fill mining and filling of large stopes that required the fill to be free standing. A less rigorous approach was required for mines with smaller stopes and in situations where the quality of the fill would not be a major factor in safety and the productivity of the mine. Echo Bay’s Lamefoot Mine was a good example of a simplistic rockfill system were cemented slurry is sprayed onto rock in the back of trucks, which then haul the fill to the stope. In the case of Lamefoot, the cement slurry itself is mixed independently in a colloidal mixer and the binder components and water are carefully measured, but the rock itself is was not measured. Generally this style fill system is mobile and relocatable to other parts of the mine (Reschke, 2000).
110 6.2.2 Composite Fill Systems
Composite fill is a subject of increasing interest (Annor et. al, 2003) and has significant potential as a means of disposal of preconcentration rejects. Mt. Isa combines rock with classified hydraulic tailings and has found that composite fills address many of the short comings of rock fill, such as segregation and the problems with the cyclical nature of the filling process (Kuganathan, 2005, Kuganathan and Niedorf, 2005). The most common systems involve delivering the materials to the stope separately, with the graded aggregate delivered via conveyor belt and cemented hydraulic fill via pipeline from a surface plant. The mixing of rock and hydraulic fill occurs during placement. At Mt . Isa’s Enterprise Mine, a rocky paste fill comprising a 3:1 mixture of rock fill to cemented hydraulic fill that was mixed together prior to placement was evaluated (Kuganthan and Sheppard, 2005). The proposed placement system utilized a conveyor system for the rocky paste fill. The use of a conveyor constituted a major disadvantage due to its high costs. BHP’s Olympic Dam Mine utilizes a composite backfill, referred to as cemented aggregate fill (CAF), composed of various mixtures of deslimed tailings, crushed mine waste rock, crushed quarried rock, cement, fly ash and water. The fill is mixed in a surface plant and loaded into semi trailer tipper trucks, which deliver the fill to the stope via boreholes from the surface (Baldwin, 2000). Barrick’s Bulyanhulu operation uses a 50/50 mix of waste rock and filtered tailings to create a high strength fill that is distributed by borehole and pipeline from surface to the stope (Landriault, et al. 2000). Several operations have introduced small quantities of aggregate to their paste and hydraulic fill systems, mostly in the 25mm range, usually as a means of disposing of fine waste rock and to save on binder costs (Bloss, 2000, Grice, 1989, Landriault, 1992). Deep South African gold mines have been working for several years on the development and operation of underground backfill plants that include the crushing of development waste (Iigner, 2001). Based on this experience the authors developed a conceptualized underground backfill plant using crushed waste. Underground backfill systems had difficultly mixing tailings and crushed waste rock and as a result Iigner recommends not blending the two materials due to the variable nature of both the tailings and the crushed waste rock.
111
6.3 Conceptual Backfill Systems
Based on the literature review, two basic backfill systems were devised. One based the simplistic rockfill plants and other a composite fill system utilizing ideas from the highly engineered rockfill and composite fill plants that were described in Sections 6.2.1 and 6.2.2.
6.3.1 Rockfill
Most rock fill systems in remote mines are fairly simple, generally consisting of a piece of mobile equipment that transports the rock component of the fill to an area where a cement slurry can be applied to the rock, creating a cemented rock fill. These systems typically require minimal capital investment and can easily be moved within the mine itself. (Reschke, 2000) The rejects would be transported through the mine via passes. This system would also have the benefit that development waste is easily incorporated into the fill system if desired. Cement would be transported to the plant by borehole from surface. It can be safely assumed that water is already available in the mine. Once the fill is mixed, mobile equipment and a network of passes would facilitate the final delivery of the fill to the stope. The ultimate desire for such a system would be a setup that could be easily moved within the mine itself, so as to minimize the amount of handling once the rejects and binder were combined.
112
Figure 6.1: Diagram of Rock Fill System
6.3.2 Composite Fill
The idealized system for disposal of rejects in mines with a hydraulic or paste fill system is one that can be distributed via boreholes and pipes (Bamber et al, 2006). A similar backfill system would be used for both a paste or hydraulic fill based system. A rocky paste fill system would allow for thickened tailings to be transported from surface as slurry then mixed with rejects underground, generating a competent fill, which ultimately would require less binder for a given strength. By mixing the tailings and cement on surface, underground development is minimized. This would also allow for more precise mixing of the binder and tailings prior to the coarser material being added. The time between the mixing of the binder and tailings on surface and the addition of the rejects underground needs to be minimized to ensure that the tailings binder mixture does not become too stiff for mixing the coarse aggregate. To ensure the appropriate size for pipe transportation, the rejects would need to be in the 37.5 mm size range as either part of the initial preconcentration step or the waste handling step. Where the final sizing of the rejects occurs would depend upon the metallurgical requirements of preconcentration. The crushed rejects and tailings would be mixed as close to the top of the ore body as possible to allow for gravity driven pipe flow and thus avoid handling by mobile equipment. The addition of rejects to a paste fill system would not be overly problematic since paste systems already work in laminar flow situations. Hydraulic fill systems
113 operate under turbulent flow conditions, which could make the integration of the rejects more troublesome. If rejects are transported by turbulent flow there is a strong probability of unacceptable wear rates with standard hardened steel pipes. Utilizing pipelines made of different materials may alleviate the wear issues. If a laminar flow is desired after the rejects are added to an hydraulic fill then care must be taken to ensure enough fine material is included in the mix design.
Figure 6. 2: Diagram of Composite Fill System
6.3.3 Linking Preconcentration Systems and Backfill Systems
In chapter three the basic operations of a preconcentration plant were stated to be: 1. Feed preparation by screening, sizing, and washing, 2. Particle separation, 3. Handling of concentrate and rejects. The most important consideration in the linkage of the preconcentration system and the backfill system disposing of the rejects is the size distribution of the material in the process and how the size distribution is managed. In a unified process of preconcentration and backfilling the size distribution will be controlled by two factors: the particle size limitations of the preconcentration system and the top size limits of the
114 backfill method. For a preconcentration system to be effective the run of mine ore must be reduced to generate an appropriate degree of liberation between the mineralized and barren particles. Additionally, particle separation technologies have a limited size range in which they are efficient as discussed in chapter three. For backfills the particle size distribution and top size is governed by concerns with segregation during placement, strength development, and top size limits imposed by transportation, such as those for pipeline reticulated fills. How the integration of the size reduction for the two separate process is done is the process that will determine the success of linking the two processes. The two basic options for size reduction are to have all size reduction done prior to particle separation or to have a distinct crushing stage for the rejects.
6.3.3.1 Single Crushing Stage for all of Preconcentration
The primary benefit of conducting all size reduction prior to particle separation is a simplified flow sheet that minimizes the amount of equipment required. Since most particle separation systems do not operate effectively on finer particles, the fines produced during crushing prior to preconcentration usually report directly to the concentrate. As a result of the scalping of fines, rejects produced from a unified system with only a single crushing stage will have a uniform coarse size distribution with almost no fines. Meaning that unless other materials; such as flotation tailings, development waste, or natural aggregates; are combined with the preconcentration rejects the fill generated will have a high void ratio, resulting in a stiff fill. Such a fill may not be appropriate for some geotechnical situations and would require mobile equipment or conveyors to transport.
115
Figure 6. 3: Single Crushing Stage Preconcentration Flow Sheet
A single crushing stage for both particle separation and backfill mixing would be ideal for smaller remote mines that have a desire to minimize equipment and personnel requirements.
6.3.3.2 Crushing Stages for Particle Separation and Reject Disposal
In cases were the size distribution requirements for the particle separation and backfill are significantly different, it maybe desired to have a separate stage of crushing for the rejects. A prime example would be a preconcentration system utilizing an automated sorter on coarse ores, where the rejects would be incorporated into a pipeline reticulated backfill. The prime advantage of having a separate crusher is that the size distribution of the rejects can be tailored specifically to fit the backfill requirements. Any fines generated during crushing will be retained in the fill, which provides for a wider size distribution and the benefits that come with it. The primary disadvantage is the capital, maintenance, and operating requirements of a separate crusher, additionally there will be
116 additional space requirements, which is an expensive and scarce commodity in the underground environment.
Figure 6. 4: Independent Crushing Stage for Particle Separation and Reject Disposal Preconcentration Flow Sheet
The additional cost of crushing the rejects separately will likely be at larger mines where high tonnage throughputs would be hindered by the additional sizing requirements of a single crushing stage. Additionally larger mines are more likely to expend the additional capital costs required by a pipeline reticulated fill, so disposing of preconcentration rejects in the same manner would be make the additional cost of a second crusher worthwhile. Smaller mines that require strong competent fills for underhand cut and fill or large free standing fills, may also find the additional costs of a second crusher acceptable.
6.4 Case Studies
The four Xstrata Nickel mines examined in the study had unique geographical locations, mining methods, and ore geologies. By looking at current practices, the test results from
117 this thesis, and observation during a site visit to each of the four mines; a basic plan for how preconcentration might be implemented at each operation is presented.
6.4.1 Thayer Lindsley
Thayer Lindsley is the smallest of Xstrata Nickel’s Sudbury mines with an annual production of 500,000 tons per annum through 2012. Run of mine ore is crushed on surface and trucked 90 km to the Strathcona mill. Currently, Thayer Lindsley utilizes long hole open stoping and cut and fill methods with uncemented rockfill. Development waste and externally sources 90mm rock provide the raw materials for the rockfill. The fill is used primarily to provide confinement for the post pillars in cut and fill mining and to maintain the overall stability of the mine. Operations plans for Thayer Lindsley have the operation transitioning to a cemented rockfill to facilitate pillar removal. For Thayer Lindsley the greatest benefit to preconcentration would be realized by reducing the total amount of ore that must be transported from the mine. Preconcentration preformed underground would also allow a higher mining rate, since the production rate is currently limited by the hoisting capacity. Since Thayer Lindsley currently has to purchase additional aggregate to meet its backfilling requirements, the utilization of the rejects would allow for a reduction in the overall minefill costs.
Geologically the ore consists of three zones, two contact ores and a footwall ore. The footwall ore sample had significantly higher copper and precious metal grades than those predicted by geologic reserves for the Thayer Lindsley, geological models portray the Thayer Lindsley footwall ore as similar to a contact ore overall with increased copper and precious metals. Inspection of the ores during a site visit indicated similarities between Thayer Lindsley footwall ores and contact ores. During preconcentration testing the Thayer Lindsley footwall ores showed results were similar to contact ore for the nickel recoveries and over mass rejection.
The dense media separation had better metallurgical results than the conductivity sensor overall (Table 6.1). However the conductivity sensor showed results comparable to the dense media separation for zone 2 and footwall ores. If a dry sorting based system
118 capable of metallurgical performance similar to a dense media separator was available it would likely be more advantageous than a dense media separation system. The relatively low mining rates (approximately 1000 tons / day) at Thayer Lindsley are well within the range of current sorter technologies. The rejects from the three ore zones had significantly different specific gravities, as a result specific gravity based separation may not achieve the best metallurgical results for separating a blend of all three ores zones simultaneously. The minefill requirements for a cement rockfill at Thayer Lindsley were stated to be approximately 1 MPa, which based on the tests conducted as part of this thesis would be easily achieved.
Table 6. 1: Preconcentration System Summary for Thayer Lindsley Orebody Ore Type Dense Media Separation Conductivity Sorter Reject Disposal Mass Mass Recovery Increase in Recovery Increase in Rejection Rejection UCS (MPa) (%) Grade (%) (%) Grade (%) (%) (%) 100% Max. Max. Ni Cu Ni Cu Ni Cu Ni Cu Reject Full CT. Zone 1 Contact 20 95 93 0.13 0.06 56 63 48 0.30 0.05 1.2 NA 1.7 Zone 2 Contact 26 98 96 0.41 0.25 38 90 84 0.63 0.00 1.6 1.8 1.4 Footwall Footwall 37 98 98 0.64 3.80 34 95 88 0.56 2.97 2.1 0.6 0.6
A preconcentration system either underground or surface with a single crushing stage prior to an automated sorter would be the recommendation for Thayer Lindsley, since such a system would be relatively simple and could be easily integrated into the current mining process. A conservative assumption of 25% mass rejection results in 125,000 tons of rejects. A surface preconcentration plant would capture all of the benefits from the backfilling and haulage from mine to mill. The additional cost of placing the preconcentration plant underground would need to be recouped solely by the increased mining rate that results from the additional hoisting capacity.
119 Surface Haulage
71% 5%
95% Sizing Sorter Aggregate
Surface
Underground 24% 100%
Stope Rockfill
Figure 6. 5: Block Diagram of Preconcentration with Mass Flow Based on Run of Mine Ore for Thayer Lindsley
6.4.2 Montcalm
Mining at Montcalm is done using open stoping methods and produces approximately 850,000 tons of ore per year with a four year mine life. The Montcalm mine is the only mine in this study not currently using some form of backfill. Ore is crushed on the surface and then hauled 100 km to the Kidd Creek Metallurgical complex for processing. For Montcalm, the reduction in haulage costs would be the primary motivation for preconcentration. Being a shallow mine in very competent rock, waste management provides the only reason backfilling would occur at Montcalm.
Geologically there are two main ore zones both composed of disseminated nickel and copper sulfides with the only noticeable difference between the two being the metal grades. Unlike the other mines in the study there were no economic levels of precious metals.
Dense media separation showed very positive metallurgical results for the Montcalm deposit (Table 6.2). The rejects from both ore zones had essentially the same specific
120 gravity, making processing ore from the two zones simultaneously feasible. A significant water containment and treatment system already exists at the mine that could accommodate any effluent generated by a dense media system.
Table 6. 2: Preconcentration System Summary for Montcalm Orebody Ore Type Dense Media Separation Conductivity Sorter Reject Disposal Mass Mass Recovery Increase in Recovery Increase in Rejection Rejection UCS (MPa) (%) Grade (%) (%) Grade (%) (%) (%) 100% Max. Max. Ni Cu Ni Cu Ni Cu Ni Cu Reject Full CT. East Disseminated 25 98 93 0.50 0.16 25 94 85 0.40 0.07 1.4 1.2 1.0 West Disseminated 32 98 95 0.13 0.05 70 59 58 0.32 0.15 3.4 1.2 0.9
Based on the dense media separation results, a mass rejection of 25%is a reasonable expectation. Such a rejection would produce 213,000 tons of DMS rejects a year. The large stopes could be easily backfilled by mobile equipment or waste passes from surface, with dilution being the only concern. If dilution did prove to be problematic, then some cement might be used to stabilize the rejects returned to the mined out stopes. Surface preconcentration would be best suited for Montcalm, since there would be no real benefit derive from the additional cost to put a plant underground (Figure 6.6). A crusher that reduces the ore to a size appropriate for dense media separation is already in use at Montcalm; as such the only additional piece of equipment outside of the dense media separator itself that is required would be a screen to remove fines.
121 Surface Haulage
10% 68 %
90% Sizing DMS
Surface
Underground 100% 22%
Stope Underground Storage
Figure 6. 6: Block Diagram of Preconcentration with Mass Flow Based on Run of Mine for Montcalm
6.4.3 Craig
Located near the Strathcona mill in Onaping, the Craig Mine has the highest mining rate of the Sudbury operations at 800,000 tons per annum. The Craig mine uses a combination of long hole open stoping and cut and fill mining. Hydraulic classified fill, both cemented and uncemented, is the primary backfilling method utilized. Classified flotation tails, delivered from the Strathcona Mill by truck are mixed in a surface fill plant with water and binder before being sent underground via pipeline. Due to the short distance between Craig mine and Strathcona mill, preconcentration would need to occur underground to capture the maximum value.
The two deposits examined at Craig were both contact ores with similar characteristics.
Overall the ores from both ore zones responded favorably to both dense media and conductivity sorting (Table 6.3). In general a sorter would be a preferred choice for the
122 underground environment, since it is a dry process and with the exception of the sophisticated electronics all the components in the system are already common place underground.
Table 6. 3: Preconcentration System Summary for C raig Orebody Ore Type Dense Media Separation Conductivity Sorter Reject Disposal Mass Mass Recovery Increase in Recovery Increase in Rejection Rejection UCS (MPa) (%) Grade (%) (%) Grade (%) (%) (%) 100% Max. Max. Ni Cu Ni Cu Ni Cu Ni Cu Reject Full CT. 8112 Contact 14 98 98 0.14 0.06 28 93 87 0.34 0.10 NA 0.8 1.6 LGBX Contact 32 97 82 1.06 0.07 17 96 87 0.33 0.02 1.5 0.5 1.4
If it is assumed that a preconcentration system will result in 20 25% mass rejection from run of mine ore, approximately 160,000 200,000 tons of reject material a year will be produced. A composite backfill system would allow for the use of the current distribution network and produce a 1.5 MPa fill with minimal binder addition. A preconcentration system consisting of independent crushing for the particle separation and reject disposal would be justified, by the limited top size of the rejects required by a pipeline reticulated fill (Figure 6.7).
123
Flotation Fill Plant Plant
10% 72% Surface Underground Sizing Sorter 90% 100% 18%
Composite Stope Fill Plant with Crusher
Figure 6. 7: Block Diagram of Preconcentration with Mass Flow Based on Run of Mine for Craig
6.4.4 Fraser Mine
Fraser Mine is the only Xstrata mine in the study where the mine was able to feed the mill directly from underground without the need for additional haulage on surface. Mining at Fraser is carried out using long hole open stoping and cut and fill methods. Backfill needs are met with hydraulic fill produced at the Strathcona Mill and reticulated underground via pipeline. Two distinct orebodies are being mined at the Fraser mine; each showing markedly different geologic characteristics resulting in different mining methods and metallurgical responses (Table 6.4). Fraser Nickel is a contact ore similar to those at the Craig Mine which is mined at a rate of 500,000 tons per annum. The Fraser Copper ores are a narrow vein footwall ore with barren host rock, mined at 250,000 tons per annum. Fraser Copper is some of the most highly stressed ground witnessed in this study and requires a competent fill in order to be mined. The difference between Fraser Nickel and Fraser Copper would justify two separate preconcentration systems.
124 Table 6. 4: Preconcentration System Summary for Fraser Mine Orebody Ore Type Dense Media Separation Conductivity Sorter Reject Disposal Mass Mass Recovery Increase in Recovery Increase in Rejection Rejection UCS (MPa) (%) Grade (%) (%) Grade (%) (%) (%) 100% Max. Max. Ni Cu Ni Cu Ni Cu Ni Cu Reject Full CT. Nickel Contact 25 91 92 0.14 0.08 20 93 89 0.13 0.04 1.8 1.2 1.4 Copper Footwall 53 96 98 0.43 11.5 59 81 75 0.82 9.50 2.1 1.3 1.3
While preconcentration would have some value at Fraser Nickel, the proximity of Fraser Nickel to the surface plant and the metallurgical performance of the ore make preconcentration of the Fraser Nickel orebody the least attractive of nine orebodies in this study. Preconcentration at Fraser Nickel would require a system similar to that prescribed for the Craig Mine (Figure 6.7). Such a system can be assumed to reject 20 25% of the overall run of mine material producing 100,000 to 125,000 tons per annum of rejects. A composite fill of flotation tailings and Fraser Nickel rejects could be expected to generate a fill consistently in excess of 1 MPa.
Fraser Copper had the best metallurgical performance for preconcentration, with the resulting concentrate possibly being smeltable (Table 6.4). Due to mineralogy, Fraser Copper ores would be amenable to almost all forms of sorting due to the clear difference between barren and mineralized material, even though it wasn’t clearly evident in the conductivity results shown in this thesis. As an underground operation, a sorter would likely be the preferred choice. The high rejection rates of preconcentration for Fraser Copper, often in excess of 50%, would produce 125,000 tons or more of rejects. When the addition of development waste is considered this would likely come close to meeting the backfill needs for Fraser Copper. While the rejects and development waste alone would produce a rockfill able to meet the backfill requirements for Fraser Copper, the addition of flotation tailings to create a composite fill is still recommended. A composite fill would provide a more resilient fill that provides better regional support in the high stress environment on Fraser Copper. The high stresses found a Fraser Copper and the high grade ores that are recoverable with a competent fill would justify a sophisticated fill system. The desire for a competent fill and the minimal size reduction required to achieve liberation for the particle separation stage make a separate crusher for the rejects
125 beneficial to Fraser Copper. A separate crusher would allow for more control of the reject size distribution and could be designed to accept development waste, in addition the fines produced by the crusher would be included in the fill. The end result for Fraser Copper would be relatively simple particle separation stage tied with a more sophisticated composite fill system (Figure 6.8).
Smelter Fill Plant
20% 40% Surface Underground Sizing Sorter 80% 100 40%
Composite Stope Fill Plant with Crusher
Development Waste
Figure 6. 8: Block Diagram of Preconcentration with Mass Flow Based on Run of Mine for Fraser Copper
126
6.5 Conclusions and Recommendations
When evaluating the proper backfill system for a given preconcentration system a few keys points need to be considered:
1. Quality and consistency of fill required by mining method, 2. Materials available for backfill at a given operation, 3. Capital requirements of a given fill system, 4. Existing backfill systems, 5. Appropriate distribution system, 6. Distance between mine and concentrator.
127 Chapter 7 Conclusions and Recommendation
7.1 Conclusions
Several important conclusions can be drawn from the work presented in this thesis. Overall the most significant conclusion and contribution from this work is showing that preconcentration is a possibility for a wide range of nickel sulfide orebodies and that the preconcentrate rejects of these orebodies can be effectively used as an aggregate in backfill mixes or for other uses in the mining environment.
7.1.2 Conclusions from Metallurgical Work
Based on the metallurgical work done with the nine Xstrata Nickel ore samples the following can be concluded:
1. Based on the dense media separation and conductivity sorter results all nine of the orebodies are amenable to preconcentration, with the majority of orebodies showing 95% + metal recoveries and 25 % + mass rejections. 2. Orebodies of similar mineralogy had similar responses to preconcentration tests. Contact orebodies had nickel recoveries of 91 97 % and mass rejections of 20 – 25 %. Footwall ores had nickel recoveries of 97 % and mass rejections of 38 – 53 %. 3. Mass rejection for contact or massive disseminated orebodies was largely a function of the mineralogy of the ores, as opposed to footwall orebodies were mass rejection was a function of the mining dilution. 4. Conductivity sorting results showed results that were comparable to the dense media separation results, indicating that sorting is a valid process for preconcentration. 5. Conductivity sorter results were largely a function of the particle size and overall grade of a deposit. 6. The samples from the nine orebodies were not representation of the overall size distribution expect in the run of mine muck pile. The finest and coarsest
128 fractions were under represented in the nine samples. This was a result of sampling method and sample size.
7.1.3 Conclusions from Geotechnical Characterization of Rejects
The geotechnical characterization of the rejects allowed for several basic conclusions:
1. Based on the criteria evaluated in this characterization, the preconcentration rejects could be utilized in a backfill. 2. Most orebodies exhibited fairly high void ratios as a result of their fairly narrow size distributions. 3. The rejects from the nine orebodies in this study had a tendency to generate flat and elongated particles, which could negatively impact the strength and packing density of these rejects. 4. The size distribution of the rejects was the property that had the most effect on how the rejects could be used in a backfill. 5. Due to the sulfides present in the rejects acid generation is a strong possibility that would need to be considered when disposing of preconcentration rejects. Overall the use of preconcentration rejects as an aggregate within the mine environment is a prudent means of maximizing the value of a given orebody.
7.1.4 Conclusions from Fill Mix Testing
The fill testing allowed for several conclusions regarding the properties of fills composed of rejects and the overall applicability of rejects for use as minefills:
1. Rejects can be used to produce rockfills and composite minefills that equal or exceed the performance of fills currently in use. 2. There is a fundamental difference in strength development of rockfills and composite fills. Rockfills exhibit stiff and brittle characteristics, while composite fills are more plastic and yielding characteristics.
129 3. Composite fills have the overall strength characteristics of a paste or hydraulic fill regardless of reject content, until the reject content is high enough for the rejects to come into point to point contact. 4. Composite fills with high reject contents are highly sensitive to container wall effect. As result the UCS results found in this work likely did not arrive at the actually peak UCS value for a given mix, due to the cylinder size used in the testing.
7.1.5 Conclusions from Conceptual Design of Waste Handling Systems
The conceptual design work brought several conclusions:
1. The type of waste system needs to fit with the mines existing infrastructure. 2. The waste handling system should be considered as an integral part of the preconcentration systems particle sizing process.
7.2 Recommendations
Future work should be able to use the results and data generated in this work to move beyond a basic amenability level of work into a feasibility level for a given orebody. Recommendations for future work are:
1. The feed size for the preconcentration step should consider the needs of not only the particle separation, but also the reject disposal. In order to minimize the amount of communition required underground the two processes should be consider together. Traditionally the fills have been designed with the assumption that the size distribution of metallurgical waste is not something the fill engineer has control over, in preconcentration this notion can and should be challenged.
2. A large sample from a single source should be used for the next level of testing for both metallurgical and fill work. The number of variables that
130 need to be further refined will require a large volume of test work. The size of the material feed to the preconcentration system needs to be explored. Also further fill work will need to be done with the appropriate cylinder size of 6 to 10 times the maximum particle size resulting in a single cylinder requiring 8 to 60 kilograms of rejects.
3. Mechanical properties; such as triaxial strength. Modulus of Elasticity, compressibility, and permeability, require further study. The stope environment in which the fill will be disposed of should also be considered when determining the desired characteristics of the fill.
4. Future work in fills should work to determine the trade offs between rheology and strength of composite fills. There is little doubt that high strength fills composed of coarse aggregates and fine tailings can be created, it is the transportation and placement of such fills that will pose the largest technical and economic hurdle. In order for such a system to be a success both concerns need to be address with equal weight.
5. While the technical data for the metallurgical preconcentration and fill work has been examined on a fairly in depth basis, there is still lacking a significant amount of work on the mining engineering side of the underground preconcentration concept. Using the knowledge that has been generated with this and previous research an indepth study into mine planning and geotechnical concerns should be conducted. Part of this work should be the determination of the appropriate fill strengths and filling schedule required.
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139 Appendix 1 – Grades for Xstrata Nickel Samples
Craig 8112 Source Ni Cu Co Au Ag Pt Pd % % % g/t g/t g/t g/t Geological Size Assay 1.273 0.508 0.041 0.139 1.662 0.178 0.192 Metalurgical 1.113 0.480 0.033 0.079 1.419 0.095 0.104 Deviation 0.11 0.02 0.01 0.04 0.17 0.06 0.06
Craig LGBX Source Ni Cu Co Au Ag Pt Pd % % % g/t g/t g/t g/t Geological 1.790 0.410 0.057 0.038 1.884 0.283 0.251 Size Assay 1.922 0.434 0.057 0.015 1.139 0.168 0.226 Metalurgical 2.281 0.326 0.060 0.020 0.804 0.105 0.109 Deviation 0.25 0.06 0.00 0.01 0.55 0.09 0.08
Fraser Ni Source Ni Cu Co Au Ag Pt Pd % % % g/t g/t g/t g/t Geological 0.870 0.140 Size Assay 0.669 0.279 0.026 0.129 1.069 0.090 0.065 Metalurgical 0.743 0.378 0.024 0.047 1.219 0.120 0.077 Deviation 0.10 0.12 0.00 0.06 0.11 0.02 0.01
Fraser Cu Source Ni Cu Co Au Ag Pt Pd % % % g/t g/t g/t g/t Geological* 1.980 27.830 3.940 133.700 5.620 7.300 Size Assay 0.753 10.262 0.009 0.517 47.105 1.543 2.109 Metalurgical 0.614 10.939 0.008 0.138 40.355 1.813 1.748 Deviation 0.75 9.95 0.00 2.09 52.05 2.28 3.11 * assumes no dilution
140 TL 15 2 Source Ni Cu Co Au Ag Pt Pd % % % g/t g/t g/t g/t Geological 1.280 1.920 0.051 0.310 4.610 2.130 1.780 Size Assay 1.578 8.351 0.074 0.307 22.904 2.300 2.366 Metalurgical 1.245 8.136 0.052 0.661 17.515 1.099 2.180 Deviation 0.18 3.65 0.01 0.20 9.40 0.65 0.30
TL 80 Source Ni Cu Co Au Ag Pt Pd % % % g/t g/t g/t g/t Geological 1.380 0.310 0.051 0.200 4.910 0.591 0.565 Size Assay 0.733 0.513 0.026 0.140 2.015 0.224 0.192 Metalurgical 1.349 0.755 0.043 0.119 3.457 0.515 0.278 Deviation 0.37 0.22 0.01 0.04 1.45 0.19 0.20
TL 670 Source Ni Cu Co Au Ag Pt Pd % % % g/t g/t g/t g/t Geological 1.010 0.580 0.044 0.050 2.530 0.202 0.252 Size Assay 0.582 0.362 0.028 0.039 1.211 0.171 0.143 Metalurgical 0.685 0.411 0.029 0.084 1.277 0.093 0.112 Deviation 0.22 0.11 0.01 0.02 0.74 0.06 0.07
Montcalm Low Source Ni Cu Co Au Ag Pt Pd % % % g/t g/t g/t g/t Geological 1.245 0.625 0.049 Size Assay 0.829 0.310 0.031 0.022 1.118 0.003 0.008 Metalurgical 0.369 0.169 0.015 0.027 0.806 0.000 0.000 Deviation 0.44 0.23 0.02 0.00 0.22 0.00 0.01
Montcalm High Source Ni Cu Co Au Ag Pt Pd % % % g/t g/t g/t g/t Geological 1.150 0.500 0.030 Size Assay 1.118 0.842 0.058 0.057 2.678 0.013 0.011 Metalurgical 1.637 0.610 0.053 0.069 1.695 0.004 0.003 Deviation 0.29 0.17 0.01 0.01 0.69 0.01 0.01
141 Appendix 2 – Metallurgical Balances
142 Summary
Deposit DMS Separation Weight Grade (%) Distribution (%) Products SG (g) (%) Ni Cu Mg Ni Cu Mg Craig 8112 With Fines Concentrate +2.95 34639.30 86.19 1.263 0.57 5.710 97.63 96.74 86.891 Waste 2.95 5548.20 13.81 0.191 0.12 5.378 2.37 2.37 13.109 Total 40187.50 100.00 1.115 0.51 5.664 100.00 100.00 100.000 Without Fines Concentrate +2.95 31912.50 85.19 1.217 0.57 5.799 97.34 96.45 86.115 Waste 2.95 5548.20 14.81 0.191 0.12 5.378 2.66 2.66 13.885 Total 37460.70 100.00 1.065 0.50 5.737 100.00 100.00 100.000 Craig LGBX With Fines Concentrate +2.95 34571.40 67.88 3.518 0.38 2.368 97.18 81.55 67.828 Waste 2.95 16359.50 32.12 0.216 0.18 2.373 2.82 18.45 32.172 Total 50930.90 100.00 2.457 0.31 2.370 100.00 100.00 100.000 Without Fines Concentrate +2.95 31202.90 65.60 3.533 0.37 2.416 96.90 79.75 66.003 Waste 2.95 16359.50 34.40 0.216 0.18 2.373 3.10 20.25 33.997 Total 47562.40 100.00 2.392 0.31 2.401 100.00 100.00 100.000 Fraser Ni With Fines Concentrate +2.9 18938.79 51.78 1.085 0.64 3.504 82.83 83.12 42.703 Waste 2.9 17639.70 48.22 0.242 0.14 5.047 17.17 16.88 57.297 Total 36578.49 100.00 0.678 0.40 4.248 100.00 100.00 100.000 Without Fines Concentrate +2.9 15602.82 46.94 1.137 0.67 3.447 80.64 80.94 37.656 Waste 2.9 17639.70 53.06 0.242 0.14 5.047 19.36 19.06 62.344 Total 33242.52 100.00 0.662 0.39 4.296 100.00 100.00 100.000 Fraser Cu With Fines Concentrate +2.9 18592.30 46.63 0.837 22.01 0.686 96.01 97.95 17.491 Waste 2.9 21276.92 53.37 0.030 0.40 2.826 3.99 2.05 82.509 Total 39869.22 100.00 0.407 10.48 1.828 100.00 100.00 100.000 Without Fines Concentrate +2.9 11860.10 35.79 0.697 23.62 0.538 92.74 97.03 9.600 Waste 2.9 21276.92 64.21 0.030 0.40 2.826 7.26 2.97 90.400 Total 33137.02 100.00 0.269 8.71 2.007 100.00 100.00 100.000 TL-15-2 With Fines Concentrate +2.9 22152.10 63.40 1.828 10.79 1.027 97.65 97.88 43.684 Waste 2.9 12789.50 36.60 0.076 0.40 2.292 2.35 2.12 56.316 Total 34941.60 100.00 1.187 6.99 1.490 100.00 100.00 100.000 Without Fines Concentrate +2.9 19752.10 60.70 1.895 11.11 0.946 97.47 97.70 38.937 Waste 2.9 12789.50 39.30 0.076 0.40 2.292 2.53 2.30 61.063 Total 32541.60 100.00 1.180 6.90 1.475 100.00 100.00 100.000 TL-80 With Fines Concentrate +2.9 30834.60 74.27 1.702 1.11 3.579 97.73 95.65 71.138 Waste 2.9 10680.70 25.73 0.114 0.15 4.192 2.27 4.35 28.862 Total 41515.30 100.00 1.293 0.86 3.737 100.00 100.00 100.000 Without Fines Concentrate +2.9 26934.60 71.61 1.706 1.08 3.635 97.42 94.93 68.623 Waste 2.9 10680.70 28.39 0.114 0.15 4.192 2.58 5.07 31.377 Total 37615.30 100.00 1.254 0.82 3.793 100.00 100.00 100.000 TL-670 With Fines Concentrate +2.9 33294.38 80.48 0.820 0.45 6.196 95.40 92.60 80.212 Waste 2.9 8076.20 19.52 0.163 0.15 6.301 4.60 7.40 19.788 Total 41370.58 100.00 0.692 0.39 6.216 100.00 100.00 100.000 Without Fines Concentrate +2.9 28994.38 78.21 0.713 0.41 6.448 94.02 90.91 78.605 Waste 2.9 8076.20 21.79 0.163 0.15 6.301 5.98 9.09 21.395 Total 37070.58 100.00 0.593 0.36 6.416 100.00 100.00 100.000 Mont High With Fines Concentrate +2.95 33308.70 74.50 2.117 0.82 4.393 97.56 93.11 67.417 Waste 2.95 11398.30 25.50 0.154 0.18 6.205 2.44 6.89 32.583 Total 44707.00 100.00 1.617 0.66 4.855 100.00 100.00 100.000 Without Fines Concentrate +2.95 30303.70 72.67 2.144 0.84 4.383 97.36 92.62 65.253 Waste 2.95 11398.30 27.33 0.154 0.18 6.205 2.64 7.38 34.747 Total 41702.00 100.00 1.600 0.66 4.881 100.00 100.00 100.000 Mont Low With Fines Concentrate +2.95 10945.60 28.73 1.173 0.45 5.500 79.51 69.47 28.725 Waste 2.95 27148.30 71.27 0.122 0.08 5.502 20.49 30.53 71.275 Total 38093.90 100.00 0.424 0.19 5.501 100.00 100.00 100.000 Without Fines Concentrate +2.95 6780.60 19.98 1.585 0.59 5.357 76.46 65.00 19.563 Waste 2.95 27148.30 80.02 0.122 0.08 5.502 23.54 35.00 80.437
143 Craig 8112
DMS Concentrates Product Separation SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg With Fines Concentrate +2.95 34639.30 86.19 1.263 0.57 5.710 97.63 96.74 86.891 Waste 2.95 5548.20 13.81 0.191 0.12 5.378 2.37 3.26 13.109 Total 40187.50 100.00 1.115 0.51 5.664 100.00 100.00 100.000 Without Fines Concentrate +2.95 31912.50 85.19 1.217 0.57 5.799 97.34 96.45 86.115 Waste 2.95 5548.20 14.81 0.191 0.12 5.378 2.66 3.55 13.885 Total 37460.70 100.00 1.065 0.50 5.737 100.00 100.00 100.000
Calculations
Crushed Material SG Size Fraction Weight Grade (%) Distribution (%) (mm) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 75+25 5591.40 79.37 1.717 1.05 4.803 73.814 74.26 81.715 25+6.7 1453.60 20.63 2.343 1.40 4.134 26.186 25.74 18.285 +3.1 total 7045.00 100.00 1.846 1.12 4.665 100.000 100.00 100.000 3.1+2.95 75+25 7675.00 76.94 0.241 0.15 7.890 74.447 75.78 80.003 25+6.7 2300.30 23.06 0.276 0.16 6.580 25.553 24.22 19.997 -3.1+2.95 total 9975.30 100.00 0.249 0.15 7.588 100.000 100.00 100.000 2.95+2.8 75+25 569.40 72.48 0.073 0.05 3.499 65.115 62.21 79.303 25+6.7 216.20 27.52 0.103 0.08 2.405 34.885 37.79 20.697 -2.95+2.8 total 785.60 100.00 0.081 0.06 3.198 100.000 100.00 100.000 2.8 75+25 353.30 21.37 0.112 0.08 1.990 12.558 17.86 7.584 25+6.7 1299.70 78.63 0.212 0.10 6.592 87.442 82.14 92.416 -2.8 total 1653.00 100.00 0.191 0.10 5.608 100.000 100.00 100.000 Fines 1276.80 100.00 1.280 0.68 5.782 100.000 100.00 100.000
Crushed Material - Summary SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 7045.00 33.98 1.846 1.12 4.665 74.304 75.31 25.731 3.1+2.95 9975.30 48.11 0.249 0.15 7.588 14.194 14.47 59.263 2.95+2.8 785.60 3.79 0.081 0.06 3.198 0.365 0.44 1.967 2.8 1653.00 7.97 0.191 0.10 5.608 1.800 1.51 7.258 Fines 1276.80 6.16 1.280 0.68 5.782 9.337 8.27 5.780 Total 20735.70 100.00 0.844 0.51 6.160 100.000 100.00 100.000
Uncrushed Material SG Size Fraction Weight Grade (%) Distribution (%) (mm) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 75+25 7791.40 86.15 2.275 0.89 3.754 80.039 87.24 92.367 25+6.7 1252.30 13.85 3.530 0.81 1.930 19.961 12.76 7.633 +3.1 total 9043.70 100.00 2.449 0.88 3.501 100.000 100.00 100.000 3.1+2.95 75+25 5286.70 90.39 0.154 0.11 7.899 67.304 70.65 93.140 25+6.7 561.80 9.61 0.704 0.43 5.475 32.696 29.35 6.860 -3.1+2.95 total 5848.50 100.00 0.207 0.14 7.666 100.000 100.00 100.000 2.95+2.8 75+25 1569.00 64.80 0.208 0.17 7.116 65.913 69.09 67.232 25+6.7 852.40 35.20 0.198 0.14 6.384 34.087 30.91 32.768 -2.95+2.8 total 2421.40 100.00 0.204 0.16 6.858 100.000 100.00 100.000 2.8 75+25 488.80 71.03 0.308 0.09 1.651 80.149 55.07 55.707 25+6.7 199.40 28.97 0.187 0.18 3.218 19.851 44.93 44.293 -2.8 total 688.20 100.00 0.273 0.12 2.105 100.000 100.00 100.000 Fines 1450.00 100.00 2.253 0.57 3.685 100.000 100.00 100.000
144 Uncrushed Material - Summary SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 9043.70 46.49 2.449 0.88 3.501 81.105 78.98 31.698 3.1+2.95 5848.50 30.07 0.207 0.14 7.666 4.430 8.18 44.880 2.95+2.8 2421.40 12.45 0.204 0.16 6.858 1.813 3.84 16.623 2.8 688.20 3.54 0.273 0.12 2.105 0.688 0.79 1.450 Fines 1450.00 7.45 2.253 0.57 3.685 11.964 8.21 5.349 Total 19451.80 100.00 1.404 0.52 5.136 100.000 100.00 100.000
Crushed and Uncrushed - Combined SG Product Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 crushed 7045.00 43.79 1.846 1.12 4.665 37.000 49.865 50.929 uncrushed 9043.70 56.21 2.449 0.88 3.501 63.000 50.135 49.071 +3.1 total 16088.70 100.00 2.185 0.99 4.011 100.000 100.000 100.000 3.1+2.95 crushed 9975.30 63.04 0.249 0.15 7.588 67.255 64.860 169.914 uncrushed 5848.50 36.96 0.207 0.14 7.666 32.745 35.140 100.648 -3.1+2.95 total 15823.80 100.00 0.233 0.15 7.617 100.000 100.000 270.562 2.95+2.8 crushed 785.60 24.50 0.081 0.06 3.198 11.420 10.598 13.140 uncrushed 2421.40 75.50 0.204 0.16 6.858 88.580 89.402 86.860 -2.95+2.8 total 3207.00 100.00 0.174 0.13 5.962 100.000 100.000 100.000 2.8 crushed 1653.00 70.60 0.191 0.10 5.608 62.652 66.452 86.485 uncrushed 688.20 29.40 0.273 0.12 2.105 37.348 33.548 13.515 -2.8 total 2341.20 100.00 0.215 0.10 4.579 100.000 100.000 100.000 Fines crushed 1276.80 46.82 1.280 0.68 5.782 33.345 51.231 58.012 uncrushed 1450.00 53.18 2.253 0.57 3.685 66.655 48.769 41.988 Fines total 2726.80 100.00 1.797 0.62 4.667 100.000 100.000 100.000
With Fines SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 16088.70 46.45 2.185 0.99 4.011 80.35 79.70 32.627 3.1+2.95 15823.80 45.68 0.233 0.15 7.617 8.44 11.78 60.939 Fines 2726.80 7.87 1.797 0.62 4.667 11.20 8.52 6.434 +2.95 with fines 34639.30 100.00 1.263 0.57 5.710 100.00 100.00 100.000 2.95+2.8 3207.00 57.80 0.174 0.13 5.962 52.64 64.46 64.075 2.8 2341.20 42.20 0.215 0.10 4.579 47.36 35.54 35.925 Waste 5548.20 100.00 0.191 0.12 5.378 100.00 100.00 100.000
Without Fines SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 16088.70 50.42 2.185 0.99 4.011 90.49 87.13 34.870 3.1+2.95 15823.80 49.58 0.233 0.15 7.617 9.51 12.87 65.130 +2.95 total 31912.50 100.00 1.217 0.57 5.799 100.00 100.00 100.000 2.95+2.8 3207.00 57.80 0.174 0.13 5.962 52.64 64.46 64.075 2.8 2341.20 42.20 0.215 0.10 4.579 47.36 35.54 35.925 Waste 5548.20 100.00 0.191 0.12 5.378 100.00 100.00 100.000
DMS Concentrates Product Separation SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg With Fines Concentrate +2.95 34639.30 86.19 1.263 0.57 5.710 97.63 96.74 86.891 Waste 2.95 5548.20 13.81 0.191 0.12 5.378 2.37 3.26 13.109 Total 40187.50 100.00 1.115 0.51 5.664 100.00 100.00 100.000 Without Fines Concentrate +2.95 31912.50 85.19 1.217 0.57 5.799 97.34 96.45 86.115 Waste 2.95 5548.20 14.81 0.191 0.12 5.378 2.66 3.55 13.885 Total 37460.70 100.00 1.065 0.50 5.737 100.00 100.00 100.000
145 Craig LGBX
DMS Concentrates Product Separation SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg With Fines Concentrate +2.95 34571.40 67.88 3.518 0.38 2.368 97.18 81.55 67.828 Waste 2.95 16359.50 32.12 0.216 0.18 2.373 2.82 18.45 32.172 Total 50930.90 100.00 2.457 0.31 2.370 100.00 100.00 100.000 Without Fines Concentrate +2.95 31202.90 65.60 3.533 0.37 2.416 96.90 79.75 66.003 Waste 2.95 16359.50 34.40 0.216 0.18 2.373 3.10 20.25 33.997 Total 47562.40 100.00 2.392 0.31 2.401 100.00 100.00 100.000
Calculations
Crushed Material SG Size Fraction Weight Grade (%) Distribution (%) (mm) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 75+25 11263.10 71.76 4.319 0.40 2.239 67.929 76.63 78.820 25+6.7 4432.00 28.24 5.182 0.31 1.529 32.071 23.37 21.180 +3.1 total 15695.10 100.00 4.563 0.37 2.039 100.000 100.00 100.000 3.1+2.95 75+25 4742.70 82.66 0.191 0.14 4.909 72.237 77.84 81.327 25+6.7 994.70 17.34 0.350 0.19 5.374 27.763 22.16 18.673 -3.1+2.95 total 5737.40 100.00 0.219 0.15 4.990 100.000 100.00 100.000 2.95+2.8 75+25 3520.10 53.67 0.284 0.13 3.309 61.262 34.97 46.999 25+6.7 3039.20 46.33 0.208 0.28 4.322 38.738 65.03 53.001 -2.95+2.8 total 6559.30 100.00 0.249 0.20 3.778 100.000 100.00 100.000 2.8 75+25 4488.00 63.80 0.181 0.15 1.007 66.740 63.80 59.029 25+6.7 2546.10 36.20 0.159 0.15 1.232 33.260 36.20 40.971 -2.8 total 7034.10 100.00 0.173 0.15 1.088 100.000 100.00 100.000 Fines 2892.50 100.00 3.200 0.34 2.064 100.000 100.00 100.000
Crushed Material - Summary SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 15695.10 41.39 4.563 0.37 2.039 84.278 58.33 32.307 3.1+2.95 5737.40 15.13 0.219 0.15 4.990 1.476 8.46 28.907 2.95+2.8 6559.30 17.30 0.249 0.20 3.778 1.920 12.98 25.026 2.8 7034.10 18.55 0.173 0.15 1.088 1.432 10.47 7.731 Fines 2892.50 7.63 3.200 0.34 2.064 10.893 9.76 6.028 Total 37918.40 100.00 2.241 0.27 2.612 100.000 100.00 100.000
Uncrushed Material SG Size Fraction Weight Grade (%) Distribution (%) (mm) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 75+25 6394.90 96.01 5.523 0.61 0.383 96.212 96.26 88.127 25+6.7 266.10 3.99 5.226 0.57 1.240 3.788 3.74 11.873 +3.1 total 6661.00 100.00 5.511 0.61 0.417 100.000 100.00 100.000 3.1+2.95 75+25 2967.70 95.44 0.164 0.23 3.856 74.365 86.37 95.502 25+6.7 141.70 4.56 1.184 0.76 3.804 25.635 13.63 4.498 -3.1+2.95 total 3109.40 100.00 0.210 0.25 3.854 100.000 100.00 100.000 2.95+2.8 75+25 1416.00 88.44 0.281 0.28 3.078 87.043 87.71 87.451 25+6.7 185.10 11.56 0.320 0.30 3.379 12.957 12.29 12.549 -2.95+2.8 total 1601.10 100.00 0.286 0.28 3.113 100.000 100.00 100.000 2.8 75+25 1060.00 90.99 0.184 0.09 1.075 86.456 77.09 81.076 25+6.7 105.00 9.01 0.291 0.27 2.533 13.544 22.91 18.924 -2.8 total 1165.00 100.00 0.194 0.11 1.206 100.000 100.00 100.000 Fines 476.00 100.00 4.480 0.92 1.067 100.000 100.00 100.000
146 Uncrushed Material - Summary SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 6661.00 51.19 5.511 0.61 0.417 91.365 69.20 12.832 3.1+2.95 3109.40 23.90 0.210 0.25 3.854 1.629 13.49 55.323 2.95+2.8 1601.10 12.30 0.286 0.28 3.113 1.138 7.72 23.011 2.8 1165.00 8.95 0.194 0.11 1.206 0.561 2.11 6.489 Fines 476.00 3.66 4.480 0.92 1.067 5.307 7.48 2.345 Total 13012.50 100.00 3.088 0.45 1.664 100.000 100.00 100.000
Crushed and Uncrushed - Combined SG Product Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 crushed 15695.10 70.20 4.563 0.37 2.039 66.11 59.20 92.008 uncrushed 6661.00 29.80 5.511 0.61 0.417 33.89 40.80 7.992 +3.1 total 22356.10 100.00 4.845 0.44 1.555 100.00 100.00 100.000 3.1+2.95 crushed 5737.40 64.85 0.219 0.15 4.990 65.71 51.91 70.494 uncrushed 3109.40 35.15 0.210 0.25 3.854 34.29 48.09 29.506 -3.1+2.95 total 8846.80 100.00 0.216 0.19 4.590 100.00 100.00 100.000 2.95+2.8 crushed 6559.30 80.38 0.249 0.20 3.778 78.12 74.33 497.270 uncrushed 1601.10 19.62 0.286 0.28 3.113 21.88 25.67 100.000 -2.95+2.8 total 8160.40 100.00 0.256 0.22 3.648 100.00 100.00 597.270 2.8 crushed 7034.10 85.79 0.173 0.15 1.088 84.36 89.50 84.490 uncrushed 1165.00 14.21 0.194 0.11 1.206 15.64 10.50 15.510 -2.8 total 8199.10 100.00 0.176 0.14 1.105 100.00 100.00 100.000 Fines crushed 2892.50 85.87 3.200 0.34 2.064 81.28 69.19 92.160 uncrushed 476.00 14.13 4.480 0.92 1.067 18.72 30.81 7.840 Fines total 3368.50 100.00 3.381 0.42 1.923 100.00 100.00 100.000
With Fines SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 22356.10 64.67 4.845 0.44 1.555 89.07 76.42 42.479 3.1+2.95 8846.80 25.59 0.216 0.19 4.590 1.57 12.64 49.608 Fines 3368.50 9.74 3.381 0.42 1.923 9.36 10.94 7.913 +2.95 with fines 34571.40 100.00 3.518 0.38 2.368 100.00 100.00 100.000 2.95+2.8 8160.40 49.88 0.256 0.22 3.648 59.15 59.90 76.663 2.8 8199.10 50.12 0.176 0.14 1.105 40.85 40.10 23.337 Waste 16359.50 100.00 0.216 0.18 2.373 100.00 100.00 100.000
Without Fines SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 22356.10 71.65 4.845 0.44 1.555 98.27 85.80 46.129 3.1+2.95 8846.80 28.35 0.216 0.19 4.590 1.73 14.20 53.871 +2.95 total 31202.90 100.00 3.533 0.37 2.416 100.00 100.00 100.000 2.95+2.8 8160.40 49.88 0.256 0.22 3.648 59.15 59.90 76.663 2.8 8199.10 50.12 0.176 0.14 1.105 40.85 40.10 23.337 Waste 16359.50 100.00 0.216 0.18 2.373 100.00 100.00 100.000
DMS Concentrates Product Separation SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg With Fines Concentrate +2.95 34571.40 67.88 3.518 0.38 2.368 97.18 81.55 67.828 Waste 2.95 16359.50 32.12 0.216 0.18 2.373 2.82 18.45 32.172 Total 50930.90 100.00 2.457 0.31 2.370 100.00 100.00 100.000 Without Fines Concentrate +2.95 31202.90 65.60 3.533 0.37 2.416 96.90 79.75 66.003 Waste 2.95 16359.50 34.40 0.216 0.18 2.373 3.10 20.25 33.997 Total 47562.40 100.00 2.392 0.31 2.401 100.00 100.00 100.000
147 Fraser Nickel
DMS Concentrates Product Separation SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg With Fines Concentrate +2.9 18938.79 51.78 1.085 0.64 3.504 82.83 83.12 42.703 Waste 2.9 17639.70 48.22 0.242 0.14 5.047 17.17 16.88 57.297 Total 36578.49 100.00 0.678 0.40 4.248 100.00 100.00 100.000 Without Fines Concentrate +2.9 15602.82 46.94 1.137 0.67 3.447 80.64 80.94 37.656 Waste 2.9 17639.70 53.06 0.242 0.14 5.047 19.36 19.06 62.344 Total 33242.52 100.00 0.662 0.39 4.296 100.00 100.00 100.000
Calculations
Crushed Material SG Size Fraction Weight Grade (%) Distribution (%) (mm) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 75+25 755.50 31.78 1.979 2.80 1.625 28.457 60.27 31.348 25+6.7 1621.60 68.22 2.318 0.86 1.658 71.543 39.73 68.652 +3.1 total 2377.10 100.00 2.210 1.48 1.648 100.000 100.00 100.000 3.1+2.9 75+25 6870.00 83.26 0.706 0.37 3.927 79.819 80.35 79.602 25+6.7 1381.00 16.74 0.888 0.45 5.006 20.181 19.65 20.398 -3.1+2.9 total 8251.00 100.00 0.736 0.38 4.108 100.000 100.00 100.000 2.9+2.7 75+25 2699.20 29.55 0.200 0.15 4.260 29.448 36.39 27.917 25+6.7 6434.70 70.45 0.201 0.11 4.614 70.552 63.61 72.083 -2.9+2.7 total 9133.90 100.00 0.201 0.12 4.509 100.000 100.00 100.000 2.7 75+25 12.54 4.79 0.032 0.04 3.083 1.509 3.87 5.217 25+6.7 249.43 95.21 0.105 0.05 2.816 98.491 96.13 94.783 -2.7 total 261.97 100.00 0.102 0.05 2.829 100.000 100.00 100.000 Fines 2362.97 100.00 0.782 0.56 3.846 100.000 100.00 100.000
Crushed Material - Summary SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 2377.10 10.62 2.210 1.48 1.648 34.938 38.48 4.409 3.1+2.9 8251.00 36.86 0.736 0.38 4.108 40.407 34.68 38.155 2.9+2.7 9133.90 40.80 0.201 0.12 4.509 12.190 12.20 46.370 2.7 261.97 1.17 0.102 0.05 2.829 0.177 0.14 0.834 Fines 2362.97 10.56 0.782 0.56 3.846 12.288 14.51 10.231 Total 22386.94 100.00 0.672 0.41 3.968 100.000 100.00 100.000
Uncrushed Material SG Size Fraction Weight Grade (%) Distribution (%) (mm) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 75+25 1520.60 82.28 1.787 0.72 1.999 84.574 76.44 74.453 25+6.7 327.58 17.72 1.513 1.03 3.184 15.426 23.56 25.547 +3.1 total 1848.18 100.00 1.738 0.77 2.209 100.000 100.00 100.000 3.1+2.9 75+25 2746.40 87.84 1.081 0.79 3.520 92.821 94.22 81.335 25+6.7 380.14 12.16 0.604 0.35 5.836 7.179 5.78 18.665 -3.1+2.9 total 3126.54 100.00 1.023 0.74 3.802 100.000 100.00 100.000 2.9+2.7 75+25 6975.20 85.18 0.317 0.17 5.971 92.299 89.88 88.705 25+6.7 1213.80 14.82 0.152 0.11 4.369 7.701 10.12 11.295 -2.9+2.7 total 8189.00 100.00 0.293 0.16 5.734 100.000 100.00 100.000 2.7 75+25 19.82 36.15 0.080 0.17 2.617 34.235 54.61 33.852 25+6.7 35.01 63.85 0.087 0.08 2.895 65.765 45.39 66.148 -2.7 total 54.83 100.00 0.084 0.11 2.795 100.000 100.00 100.000 Fines 973.00 100.00 0.984 0.35 3.590 100.000 100.00 100.000
148 Uncrushed Material - Summary SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 1848.18 13.02 1.738 0.77 2.209 32.889 26.52 6.133 3.1+2.9 3126.54 22.03 1.023 0.74 3.802 32.741 42.64 17.855 2.9+2.7 8189.00 57.70 0.293 0.16 5.734 24.523 24.43 70.534 2.7 54.83 0.39 0.084 0.11 2.795 0.047 0.11 0.230 Fines 973.00 6.86 0.984 0.35 3.590 9.801 6.31 5.247 Total 14191.55 100.00 0.688 0.38 4.691 100.000 100.00 100.000
Crushed and Uncrushed - Combined SG Product Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 crushed 2377.10 56.26 2.210 1.48 1.648 62.05 71.02 48.960 uncrushed 1848.18 43.74 1.738 0.77 2.209 37.95 28.98 51.040 +3.1 total 4225.28 100.00 2.004 1.17 1.893 100.00 100.00 100.000 3.1+2.9 crushed 8251.00 72.52 0.736 0.38 4.108 65.52 57.87 74.036 uncrushed 3126.54 27.48 1.023 0.74 3.802 34.48 42.13 25.964 -3.1+2.9 total 11377.54 100.00 0.815 0.48 4.024 100.00 100.00 100.000 2.9+2.7 crushed 9133.90 52.73 0.201 0.12 4.509 43.35 45.75 46.730 uncrushed 8189.00 47.27 0.293 0.16 5.734 56.65 54.25 53.270 -2.9+2.7 total 17322.90 100.00 0.244 0.14 5.088 100.00 100.00 100.000 2.7 crushed 261.97 82.69 0.102 0.05 2.829 85.17 67.77 82.866 uncrushed 54.83 17.31 0.084 0.11 2.795 14.83 32.23 17.134 -2.7 total 316.80 100.00 0.099 0.06 2.823 100.00 100.00 100.000 Fines crushed 2362.97 70.83 0.782 0.56 3.846 65.87 79.53 72.236 uncrushed 973.00 29.17 0.984 0.35 3.590 34.13 20.47 27.764 Fines total 3335.97 100.00 0.841 0.50 3.771 100.00 100.00 100.000
With Fines SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 4225.28 22.31 2.004 1.17 1.893 41.21 40.94 12.054 3.1+2.9 11377.54 60.08 0.815 0.48 4.024 45.14 45.28 68.986 Fines 3335.97 17.61 0.841 0.50 3.771 13.65 13.78 18.959 +2.9 with fines 18938.79 100.00 1.085 0.64 3.504 100.00 100.00 100.000 2.9+2.7 17322.90 98.20 0.244 0.14 5.088 99.27 99.22 98.996 2.7 316.80 1.80 0.099 0.06 2.823 0.73 0.78 1.004 Waste 17639.70 100.00 0.242 0.14 5.047 100.00 100.00 100.000
Without Fines SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 4225.28 27.08 2.004 1.17 1.893 47.72 47.48 14.874 3.1+2.9 11377.54 72.92 0.815 0.48 4.024 52.28 52.52 85.126 +2.9 total 15602.82 100.00 1.137 0.67 3.447 100.00 100.00 100.000 2.9+2.7 17322.90 98.20 0.244 0.14 5.088 99.27 99.22 98.996 2.7 316.80 1.80 0.099 0.06 2.823 0.73 0.78 1.004 Waste 17639.70 100.00 0.242 0.14 5.047 100.00 100.00 100.000
DMS Concentrates Product Separation SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg With Fines Concentrate +2.9 18938.79 51.78 1.085 0.64 3.504 82.83 83.12 42.703 Waste 2.9 17639.70 48.22 0.242 0.14 5.047 17.17 16.88 57.297 Total 36578.49 100.00 0.678 0.40 4.248 100.00 100.00 100.000 Without Fines Concentrate +2.9 15602.82 46.94 1.137 0.67 3.447 80.64 80.94 37.656 Waste 2.9 17639.70 53.06 0.242 0.14 5.047 19.36 19.06 62.344 Total 33242.52 100.00 0.662 0.39 4.296 100.00 100.00 100.000
149 Fraser Copper
DMS Concentrates Product Separation SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg With Fines Concentrate +2.9 18592.30 46.63 0.837 22.01 0.686 96.01 97.95 17.491 Waste 2.9 21276.92 53.37 0.030 0.40 2.826 3.99 2.05 82.509 Total 39869.22 100.00 0.407 10.48 1.828 100.00 100.00 100.000 Without Fines Concentrate +2.9 11860.10 35.79 0.697 23.62 0.538 92.74 97.03 9.600 Waste 2.9 21276.92 64.21 0.030 0.40 2.826 7.26 2.97 90.400 Total 33137.02 100.00 0.269 8.71 2.007 100.00 100.00 100.000
Calculations
Crushed Material SG Size Fraction Weight Grade (%) Distribution (%) (mm) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 75+25 2091.60 49.17 0.412 23.10 0.444 48.865 44.07 82.674 25+6.7 2162.50 50.83 0.417 28.36 0.090 51.135 55.93 17.326 +3.1 total 4254.10 100.00 0.415 25.77 0.264 100.000 100.00 100.000 3.1+2.9 75+25 202.90 31.71 0.079 1.66 4.459 32.557 61.14 28.387 25+6.7 436.90 68.29 0.076 0.49 5.224 67.443 38.86 71.613 -3.1+2.9 total 639.80 100.00 0.077 0.86 4.981 100.000 100.00 100.000 2.9+2.7 75+25 8283.90 62.98 0.034 0.34 2.954 68.173 58.52 57.658 25+6.7 4870.00 37.02 0.027 0.41 3.690 31.827 41.48 42.342 -2.9+2.7 total 13153.90 100.00 0.031 0.37 3.226 100.000 100.00 100.000 2.7 75+25 1041.80 30.09 0.030 1.74 1.153 56.350 81.50 22.278 25+6.7 2421.00 69.91 0.010 0.17 1.731 43.650 18.50 77.722 -2.7 total 3462.80 100.00 0.016 0.64 1.557 100.000 100.00 100.000 Fines 2547.24 100.00 0.187 11.03 1.937 100.000 100.00 100.000
Crushed Material - Summary SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 4254.10 17.68 0.415 25.77 0.264 63.949 75.45 1.968 3.1+2.9 639.80 2.66 0.077 0.86 4.981 1.785 0.38 5.584 2.9+2.7 13153.90 54.68 0.031 0.37 3.226 14.982 3.31 74.357 2.7 3462.80 14.39 0.016 0.64 1.557 2.011 1.53 9.447 Fines 2547.24 10.59 0.187 11.03 1.937 17.273 19.33 8.644 Total 24057.84 100.00 0.115 6.04 2.373 100.000 100.00 100.000
Uncrushed Material SG Size Fraction Weight Grade (%) Distribution (%) (mm) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 75+25 3806.80 59.03 0.497 25.81 0.202 30.086 58.12 82.672 25+6.7 2642.20 40.97 1.664 26.80 0.061 69.914 41.88 17.328 +3.1 total 6449.00 100.00 0.975 26.22 0.144 100.000 100.00 100.000 3.1+2.9 75+25 403.80 78.07 0.246 0.58 2.182 61.776 27.89 76.938 25+6.7 113.40 21.93 0.542 5.34 2.329 38.224 72.11 23.062 -3.1+2.9 total 517.20 100.00 0.311 1.62 2.214 100.000 100.00 100.000 2.9+2.7 75+25 3497.80 78.00 0.040 0.34 2.704 79.308 77.50 79.002 25+6.7 986.60 22.00 0.037 0.35 2.548 20.692 22.50 20.998 -2.9+2.7 total 4484.40 100.00 0.039 0.34 2.670 100.000 100.00 100.000 2.7 75+25 175.82 100.00 0.008 0.06 1.858 100.000 100.00 100.000 25+6.7 0.00 0.00 0.000 0.00 0.000 0.000 0.00 0.000 -2.7 total 175.82 100.00 0.008 0.06 1.858 100.000 100.00 100.000 Fines 4184.96 100.00 1.632 24.13 0.341 100.000 100.00 100.000
150 Uncrushed Material - Summary SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 6449.00 40.79 0.975 26.22 0.144 46.731 62.06 5.887 3.1+2.9 517.20 3.27 0.311 1.62 2.214 1.195 0.31 7.248 2.9+2.7 4484.40 28.36 0.039 0.34 2.670 1.311 0.56 75.767 2.7 175.82 1.11 0.008 0.06 1.858 0.010 0.00 2.067 Fines 4184.96 26.47 1.632 24.13 0.341 50.753 37.07 9.032 Total 15811.38 100.00 0.851 17.23 0.999 100.000 100.00 100.000
Crushed and Uncrushed - Combined SG Product Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 crushed 4254.10 39.75 0.415 25.77 0.264 21.90 39.34 54.703 uncrushed 6449.00 60.25 0.975 26.22 0.144 78.10 60.66 45.297 +3.1 total 10703.10 100.00 0.752 26.04 0.192 100.00 100.00 100.000 3.1+2.9 crushed 639.80 55.30 0.077 0.86 4.981 23.44 39.61 73.566 uncrushed 517.20 44.70 0.311 1.62 2.214 76.56 60.39 26.434 -3.1+2.9 total 1157.00 100.00 0.182 1.20 3.744 100.00 100.00 100.000 2.9+2.7 crushed 13153.90 74.58 0.031 0.37 3.226 70.08 75.83 77.998 uncrushed 4484.40 25.42 0.039 0.34 2.670 29.92 24.17 22.002 -2.9+2.7 total 17638.30 100.00 0.033 0.36 3.085 100.00 100.00 100.000 2.7 crushed 3462.80 95.17 0.016 0.64 1.557 97.53 99.53 94.288 uncrushed 175.82 4.83 0.008 0.06 1.858 2.47 0.47 5.712 -2.7 total 3638.62 100.00 0.016 0.61 1.572 100.00 100.00 100.000 Fines crushed 2547.24 37.84 0.187 11.03 1.937 6.52 21.77 77.566 uncrushed 4184.96 62.16 1.632 24.13 0.341 93.48 78.23 22.434 Fines total 6732.20 100.00 1.085 19.17 0.945 100.00 100.00 100.000
With Fines SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 10703.10 57.57 0.752 26.04 0.192 51.72 68.11 16.109 3.1+2.9 1157.00 6.22 0.182 1.20 3.744 1.35 0.34 33.987 Fines 6732.20 36.21 1.085 19.17 0.945 46.93 31.55 49.903 +2.9 with fines 18592.30 100.00 0.837 22.01 0.686 100.00 100.00 100.000 2.9+2.7 17638.30 82.90 0.033 0.36 3.085 91.20 73.96 90.490 2.7 3638.62 17.10 0.016 0.61 1.572 8.80 26.04 9.510 Waste 21276.92 100.00 0.030 0.40 2.826 100.00 100.00 100.000
Without Fines SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 10703.10 90.24 0.75 26.04 0.192 97.46 99.50 32.157 3.1+2.9 1157.00 9.76 0.18 1.20 3.744 2.54 0.50 67.843 +2.9 total 11860.10 100.00 0.70 23.62 0.538 100.00 100.00 100.000 2.9+2.7 17638.30 82.90 0.03 0.36 3.085 91.20 73.96 90.490 2.7 3638.62 17.10 0.02 0.61 1.572 8.80 26.04 9.510 Waste 21276.92 100.00 0.03 0.40 2.826 100.00 100.00 100.000
DMS Concentrates Product Separation SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg With Fines Concentrate +2.9 18592.30 46.63 0.837 22.01 0.686 96.01 97.95 17.491 Waste 2.9 21276.92 53.37 0.030 0.40 2.826 3.99 2.05 82.509 Total 39869.22 100.00 0.407 10.48 1.828 100.00 100.00 100.000 Without Fines Concentrate +2.9 11860.10 35.79 0.697 23.62 0.538 92.74 97.03 9.600 Waste 2.9 21276.92 64.21 0.030 0.40 2.826 7.26 2.97 90.400 Total 33137.02 100.00 0.269 8.71 2.007 100.00 100.00 100.000
151 Thayer Lindsley Footwall
DMS Concentrates Product Separation SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg With Fines Concentrate +2.9 22152.10 63.40 1.828 10.79 1.027 97.65 97.88 43.684 Waste 2.9 12789.50 36.60 0.076 0.40 2.292 2.35 2.12 56.316 Total 34941.60 100.00 1.187 6.99 1.490 100.00 100.00 100.000 Without Fines Concentrate +2.9 19752.10 60.70 1.895 11.11 0.946 97.47 97.70 38.937 Waste 2.9 12789.50 39.30 0.076 0.40 2.292 2.53 2.30 61.063 Total 32541.60 100.00 1.180 6.90 1.475 100.00 100.00 100.000
Calculations
Crushed Material SG Size Fraction Weight Grade (%) Distribution (%) (mm) (g) (%) Ni Cu Mg Ni Cu Mg +3.35 75+25 8200.00 61.19 1.711 16.67 0.239 46.583 80.94 49.273 25+6.7 5200.00 38.81 3.094 6.19 0.388 53.417 19.06 50.727 +3.35 total 13400.00 100.00 2.248 12.60 0.297 100.000 100.00 100.000 3.35+3.1 75+25 685.20 41.42 0.656 2.13 3.647 36.761 27.16 47.634 25+6.7 969.00 58.58 0.798 4.04 2.835 63.239 72.84 52.366 -3.35++3.1 total 1654.20 100.00 0.739 3.25 3.171 100.000 100.00 100.000 3.1+2.9 75+25 886.50 64.15 0.223 1.58 5.264 61.106 63.57 68.310 25+6.7 495.40 35.85 0.254 1.62 4.370 38.894 36.43 31.690 -3.1+2.9 total 1381.90 100.00 0.234 1.59 4.944 100.000 100.00 100.000 -2.9+2.7 2600.00 46.43 0.065 0.89 2.994 46.814 57.937 41.846 3000.00 53.57 0.064 0.56 3.606 53.186 42.063 58.154 -2.9+2.7 total 5600.00 100.00 0.064 0.71 3.322 100.000 100.00 100.000 2.7 75+25 890.10 38.87 0.034 0.04 0.120 60.693 17.49 8.080 25+6.7 1400.00 61.13 0.014 0.12 0.868 39.307 82.51 91.920 -2.7 total 2290.10 100.00 0.022 0.09 0.577 100.000 100.00 100.000 Fines 1200.00 100.00 1.353 6.23 2.287 100.000 100.00 100.000
Crushed Material - Summary SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.35 13400.00 52.50 2.248 12.60 0.297 89.507 89.86 10.634 3.35+3.1 1654.20 6.48 0.739 3.25 3.171 3.634 2.86 14.026 3.1+2.9 1381.90 5.41 0.234 1.59 4.944 0.961 1.17 18.265 2.9+2.7 5600.00 21.94 0.064 0.71 3.322 1.073 2.13 49.737 2.7 2290.10 8.97 0.022 0.09 0.577 0.148 0.11 3.535 Fines 1200.00 4.70 1.353 6.23 2.287 4.825 3.98 7.338 Total 25526.20 100.00 1.318 7.36 1.465 100.148 100.11 103.535
Uncrushed Material SG Size Fraction Weight Grade (%) Distribution (%) (mm) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 75+25 1200.00 42.86 2.272 13.39 0.268 47.638 37.50 44.966 25+6.7 1600.00 57.14 1.873 16.74 0.246 52.362 62.50 55.034 +3.1 total 2800.00 100.00 2.044 15.30 0.255 100.000 100.00 100.000 3.1+2.9 75+25 389.70 75.52 0.075 0.04 3.892 58.97 7.16 78.851 25+6.7 126.30 24.48 0.161 1.60 3.221 41.03 92.84 21.149 -3.1+2.9 total 516.00 100.00 0.096 0.42 3.728 100.000 100.00 100.000 2.9+2.7 75+25 2800.00 92.50 0.189 0.22 2.953 97.45 74.88 96.175 25+6.7 227.10 7.50 0.061 0.91 1.448 2.55 25.12 3.825 -2.9+2.7 total 3027.10 100.00 0.179 0.27 2.840 100.000 100.00 100.000 2.7 75+25 1600.00 85.46 0.009 0.08 0.370 73.57 87.04 74.529 25+6.7 272.30 14.54 0.019 0.07 0.743 26.43 12.96 25.471 -2.7 total 1872.30 100.00 0.010 0.08 0.424 100.000 100.00 100.000 Fines 1200.00 100.00 1.197 10.03 1.085 100.00 100.000
152 Uncrushed Material - Summary SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 2800.00 29.74 2.044 15.30 0.255 73.641 76.42 5.364 3.1+2.9 516.00 5.48 0.096 0.42 3.728 0.638 0.39 14.428 2.9+2.7 3027.10 32.15 0.179 0.27 2.840 6.987 1.47 64.484 2.7 1872.30 19.89 0.010 0.08 0.424 0.252 0.26 5.958 Fines 1200.00 12.75 1.197 10.03 1.085 18.482 21.46 9.766 Total 9415.40 100.00 0.825 5.96 1.416 100.000 100.00 100.000
Crushed and Uncrushed - Combined SG Product Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.35 crushed 13400.00 100.00 2.248 12.60 0.297 100.00 100.00 100.000 uncrushed 0.00 0.00 0.000 0.00 0.000 0.00 0.00 0.000 +3.35 total 13400.00 100.00 2.248 12.60 0.297 100.00 100.00 100.000 3.35+3.1 crushed 1654.20 37.14 0.739 3.25 3.171 17.60 11.14 88.002 uncrushed 2800.00 62.86 2.044 15.30 0.255 82.40 88.86 11.998 -3.35+3.1 total 4454.20 100.00 1.559 10.83 1.338 100.00 100.00 100.000 3.1+2.9 crushed 1381.90 72.81 0.234 1.59 4.944 86.72 91.01 78.029 uncrushed 516.00 27.19 0.096 0.42 3.728 13.28 8.99 21.971 -3.1+2.9 total 1897.90 100.00 0.197 1.28 4.613 100.00 100.00 100.000 2.9+2.7 crushed 5600.00 64.91 0.064 0.71 3.322 39.93 82.92 68.392 uncrushed 3027.10 35.09 0.179 0.27 2.840 60.07 17.08 31.608 -2.9+2.7 total 8627.10 100.00 0.105 0.558 3.153 100.000 100.000 100.000 2.7 crushed 2290.10 55.02 0.022 0.09 0.577 71.81 58.06 62.467 uncrushed 1872.30 44.98 0.010 0.08 0.424 28.19 41.94 37.533 -2.7 total 4162.40 100.00 0.017 0.08 0.508 100.00 100.00 100.000 Fines crushed 1200.00 50.00 1.353 6.23 2.287 53.06 38.31 67.823 uncrushed 1200.00 50.00 1.197 10.03 1.085 46.94 61.69 32.177 Fines total 2400.00 100.00 1.275 8.13 1.686 100.00 100.00 100.000
With Fines SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.35 13400.00 60.49 2.248 12.60 0.297 74.37 70.65 17.491 3.35+3.1 4454.20 20.11 1.559 10.83 1.338 17.15 20.17 26.215 3.1+2.9 1897.90 8.57 0.197 1.28 4.613 0.92 1.01 38.500 Fines 2400.00 10.83 1.275 8.13 1.686 7.56 8.16 17.794 +2.9 with fines 22152.10 100.00 1.828 10.79 1.027 100.00 100.00 100.000 2.9+2.7 8627.10 67.45 0.105 0.56 3.153 92.87 93.21 92.781 2.7 4162.40 32.55 0.017 0.08 0.508 7.13 6.79 7.219 Waste 12789.50 100.00 0.076 0.40 2.292 100.00 100.00 100.000
Without Fines SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.35 13400.00 67.84 2.248 12.60 0.297 80.45 76.93 21.277 3.35+3.1 4454.20 22.55 1.559 10.83 1.338 18.55 21.97 31.889 3.1+2.9 1897.90 9.61 0.197 1.28 4.613 1.00 1.10 46.834 +2.9 total 19752.10 100.00 1.895 11.11 0.946 100.00 100.00 100.000 2.9+2.7 8627.10 67.45 0.105 0.56 3.153 92.87 93.21 92.781 2.7 4162.40 32.55 0.017 0.08 0.508 7.13 6.79 7.219 Waste 12789.50 100.00 0.076 0.40 2.292 100.00 100.00 100.000
DMS Concentrates Product Separation SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg With Fines Concentrate +2.9 22152.10 63.40 1.828 10.79 1.027 97.65 97.88 43.684 Waste 2.9 12789.50 36.60 0.076 0.40 2.292 2.35 2.12 56.316 Total 34941.60 100.00 1.187 6.99 1.490 100.00 100.00 100.000 Without Fines Concentrate +2.9 19752.10 60.70 1.895 11.11 0.946 97.47 97.70 38.937 Waste 2.9 12789.50 39.30 0.076 0.40 2.292 2.53 2.30 61.063 Total 32541.60 100.00 1.180 6.90 1.475 100.00 100.00 100.000
153 Thayer Lindsley Zone 1
DMS Concentrates Product Separation SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg With Fines Concentrate +2.9 33294.38 80.48 0.820 0.45 6.196 95.40 92.60 80.212 Waste 2.9 8076.20 19.52 0.163 0.15 6.301 4.60 7.40 19.788 Total 41370.58 100.00 0.692 0.39 6.216 100.00 100.00 100.000 Without Fines Concentrate +2.9 28994.38 78.21 0.713 0.41 6.448 94.02 90.91 78.605 Waste 2.9 8076.20 21.79 0.163 0.15 6.301 5.98 9.09 21.395 Total 37070.58 100.00 0.593 0.36 6.416 100.00 100.00 100.000
Calculations
Crushed Material SG Size Fraction Weight Grade (%) Distribution (%) (mm) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 75+25 6900.16 91.14 0.844 0.59 6.439 85.446 94.11 92.646 25+6.7 670.70 8.86 1.479 0.38 5.258 14.554 5.89 7.354 +3.1 total 7570.86 100.00 0.900 0.57 6.334 100.000 100.00 100.000 3.1+2.9 75+25 4883.20 69.73 0.266 0.13 7.141 63.709 58.78 66.710 25+6.7 2120.11 30.27 0.349 0.21 8.208 36.291 41.22 33.290 -3.1+2.9 total 7003.31 100.00 0.291 0.15 7.464 100.000 100.00 100.000 2.9+2.7 75+25 2685.90 61.53 0.155 0.07 6.298 65.428 42.74 59.408 25+6.7 1679.24 38.47 0.131 0.15 6.883 34.572 57.26 40.592 -2.9+2.7 total 4365.14 100.00 0.146 0.10 6.523 100.000 100.00 100.000 2.7 75+25 0.00 0.00 0.000 0.00 0.000 0.000 0.00 0.000 25+6.7 14.88 100.00 0.125 0.18 5.054 100.000 100.00 100.000 -2.7 total 14.88 100.00 0.125 0.18 5.054 100.000 100.00 100.000 Fines 1400.00 100.00 0.916 0.38 6.148 100.000 100.00 100.000
Crushed Material - Summary SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 7570.86 37.20 0.900 0.57 6.334 63.254 67.80 34.907 3.1+2.9 7003.31 34.41 0.291 0.15 7.464 18.922 16.93 38.048 2.9+2.7 4365.14 21.45 0.146 0.10 6.523 5.905 6.89 20.726 2.7 14.88 0.07 0.125 0.18 5.054 0.017 0.04 0.055 Fines 1400.00 6.88 0.916 0.38 6.148 11.902 8.34 6.265 Total 20354.19 100.00 0.529 0.31 6.750 100.000 100.00 100.000
Uncrushed Material SG Size Fraction Weight Grade (%) Distribution (%) (mm) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 75+25 5890.80 83.72 1.268 0.56 4.982 76.227 72.36 87.006 25+6.7 1145.30 16.28 2.034 1.10 3.827 23.773 27.64 12.994 +3.1 total 7036.10 100.00 1.393 0.65 4.794 100.000 100.00 100.000 3.1+2.9 75+25 6480.73 87.77 0.263 0.25 7.273 84.314 79.94 88.928 25+6.7 903.38 12.23 0.351 0.45 6.496 15.686 20.06 11.072 -3.1+2.9 total 7384.11 100.00 0.274 0.27 7.178 100.000 100.00 100.000 2.9+2.7 75+25 2139.50 58.57 0.169 0.22 5.952 53.575 63.34 57.170 25+6.7 1513.61 41.43 0.207 0.18 6.303 46.425 36.66 42.830 -2.9+2.7 total 3653.11 100.00 0.185 0.20 6.097 100.000 100.00 100.000 2.7 75+25 0.00 0.00 0.000 0.00 0.000 0.000 0.00 0.000 25+6.7 43.07 100.00 0.065 0.30 1.525 100.000 100.00 100.000 -2.7 total 43.07 100.00 0.065 0.30 1.525 100.000 100.00 100.000 Fines 2900.00 100.00 1.844 0.85 3.693 100.000 100.00 100.000
154 Thayer Lindsley Zone 2
DMS Concentrates Product Separation SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg With Fines Concentrate +2.9 30834.60 74.27 1.702 1.11 3.579 97.73 95.65 71.138 Waste 2.9 10680.70 25.73 0.114 0.15 4.192 2.27 4.35 28.862 Total 41515.30 100.00 1.293 0.86 3.737 100.00 100.00 100.000 Without Fines Concentrate +2.9 26934.60 71.61 1.706 1.08 3.635 97.42 94.93 68.623 Waste 2.9 10680.70 28.39 0.114 0.15 4.192 2.58 5.07 31.377 Total 37615.30 100.00 1.254 0.82 3.793 100.00 100.00 100.000
Calculations
Crushed Material SG Size Fraction Weight Grade (%) Distribution (%) (mm) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 75+25 9600.00 69.57 2.040 1.32 3.638 64.584 67.09 74.790 25+6.7 4200.00 30.43 2.557 1.48 2.803 35.416 32.91 25.210 +3.1 total 13800.00 100.00 2.197 1.37 3.384 100.000 100.00 100.000 3.1+2.9 75+25 3800.00 67.74 0.134 0.06 5.642 27.572 14.38 69.415 25+6.7 1810.00 32.26 0.739 0.75 5.219 72.428 85.62 30.585 -3.1+2.9 total 5610.00 100.00 0.329 0.28 5.506 100.000 100.00 100.000 2.9+2.7 75+25 4200.00 52.50 0.110 0.10 4.334 44.767 32.46 49.316 25+6.7 3800.00 47.50 0.150 0.23 4.923 55.233 67.54 50.684 -2.9+2.7 total 8000.00 100.00 0.129 0.16 4.614 100.000 100.00 100.000 2.7 75+25 18.20 2.16 0.219 0.48 2.288 9.145 10.53 3.805 25+6.7 825.00 97.84 0.048 0.09 1.276 90.855 89.47 96.195 -2.7 total 843.20 100.00 0.052 0.10 1.298 100.000 100.00 100.000 Fines 3100.00 100.00 1.643 1.22 3.220 100.000 100.00 100.000
Crushed Material - Summary SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 13800.00 44.01 2.197 1.37 3.384 79.093 73.69 37.188 3.1+2.9 5610.00 17.89 0.329 0.28 5.506 4.817 6.19 24.597 2.9+2.7 8000.00 25.52 0.129 0.16 4.614 2.692 5.05 29.394 2.7 843.20 2.69 0.052 0.10 1.298 0.114 0.32 0.871 Fines 3100.00 9.89 1.643 1.22 3.220 13.285 14.75 7.949 Total 31353.20 100.00 1.223 0.82 4.005 100.000 100.00 100.000
Uncrushed Material SG Size Fraction Weight Grade (%) Distribution (%) (mm) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 75+25 5200.00 87.29 2.104 1.24 2.658 84.453 87.02 88.551 25+6.7 757.20 12.71 2.660 1.27 2.360 15.547 12.98 11.449 +3.1 total 5957.20 100.00 2.175 1.24 2.620 100.000 100.00 100.000 3.1+2.9 75+25 1400.00 89.32 0.477 0.77 2.836 81.877 85.64 84.065 25+6.7 167.40 10.68 0.883 1.08 4.496 18.123 14.36 15.935 -3.1+2.9 total 1567.40 100.00 0.520 0.80 3.013 100.000 100.00 100.000 2.9+2.7 75+25 985.00 61.97 0.034 0.03 3.755 25.370 16.92 55.148 25+6.7 604.40 38.03 0.163 0.24 4.977 74.630 83.08 44.852 -2.9+2.7 total 1589.40 100.00 0.083 0.11 4.220 100.000 100.00 100.000 2.7 75+25 195.00 78.60 0.046 0.01 0.105 89.414 42.35 34.379 25+6.7 53.10 21.40 0.020 0.05 0.736 10.586 57.65 65.621 -2.7 total 248.10 100.00 0.040 0.02 0.240 100.000 100.00 100.000 Fines 800.00 100.00 1.806 1.62 3.068 100.000 100.00 100.000
155 Uncrushed Material - Summary SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 7036.10 33.48 1.393 0.65 4.794 54.909 46.49 28.160 3.1+2.9 7384.11 35.14 0.274 0.27 7.178 11.328 20.67 44.249 2.9+2.7 3653.11 17.38 0.185 0.20 6.097 3.782 7.58 18.596 2.7 43.07 0.20 0.065 0.30 1.525 0.016 0.13 0.055 Fines 2900.00 13.80 1.844 0.85 3.693 29.965 25.14 8.941 Total 21016.39 100.00 0.849 0.47 5.700 100.000 100.00 100.000
Crushed and Uncrushed - Combined SG Product Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 crushed 7570.86 51.83 0.900 0.57 6.334 41.02 48.69 58.707 uncrushed 7036.10 48.17 1.393 0.65 4.794 58.98 51.31 41.293 +3.1 total 14606.96 100.00 1.137 0.61 5.592 100.00 100.00 100.000 3.1+2.9 crushed 7003.31 48.68 0.291 0.15 7.464 50.21 34.76 49.653 uncrushed 7384.11 51.32 0.274 0.27 7.178 49.79 65.24 50.347 -3.1+2.9 total 14387.42 100.00 0.282 0.22 7.317 100.00 100.00 100.000 2.9+2.7 crushed 4365.14 54.44 0.146 0.10 6.523 48.53 37.18 56.108 uncrushed 3653.11 45.56 0.185 0.20 6.097 51.47 62.82 43.892 -2.9+2.7 total 8018.25 100.00 0.164 0.15 6.329 100.00 100.00 100.000 2.7 crushed 14.88 25.68 0.125 0.18 5.054 39.92 17.17 53.379 uncrushed 43.07 74.32 0.065 0.30 1.525 60.08 82.83 46.621 -2.7 total 57.95 100.00 0.080 0.27 2.431 100.00 100.00 100.000 Fines crushed 1400.00 32.56 0.916 0.38 6.148 19.34 17.75 44.558 uncrushed 2900.00 67.44 1.844 0.85 3.693 80.66 82.25 55.442 Fines total 4300.00 100.00 1.542 0.70 4.492 100.00 100.00 100.000
With Fines SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 14606.96 43.87 1.137 0.61 5.592 60.85 59.28 39.600 3.1+2.9 14387.42 43.21 0.282 0.22 7.317 14.87 20.73 51.035 Fines 4300.00 12.92 1.542 0.70 4.492 24.28 20.00 9.364 +2.9 with fines 33294.38 100.00 0.820 0.45 6.196 100.00 100.00 100.000 2.9+2.7 8018.25 99.28 0.164 0.15 6.329 99.65 98.70 99.723 2.7 57.95 0.72 0.080 0.27 2.431 0.35 1.30 0.277 Waste 8076.20 100.00 0.163 0.15 6.301 100.00 100.00 100.000
Without Fines SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 14606.96 50.38 1.137 0.61 5.592 80.36 74.09 43.692 3.1+2.9 14387.42 49.62 0.282 0.22 7.317 19.64 25.91 56.308 +2.9 total 28994.38 100.00 0.713 0.41 6.448 100.00 100.00 100.000 2.9+2.7 8018.25 99.28 0.164 0.15 6.329 99.65 98.70 99.723 2.7 57.95 0.72 0.080 0.27 2.431 0.35 1.30 0.277 Waste 8076.20 100.00 0.163 0.15 6.301 100.00 100.00 100.000
DMS Concentrates Product Separation SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg With Fines Concentrate +2.9 33294.38 80.48 0.820 0.45 6.196 95.40 92.60 80.212 Waste 2.9 8076.20 19.52 0.163 0.15 6.301 4.60 7.40 19.788 Total 41370.58 100.00 0.692 0.39 6.216 100.00 100.00 100.000 Without Fines Concentrate +2.9 28994.38 78.21 0.713 0.41 6.448 94.02 90.91 78.605 Waste 2.9 8076.20 21.79 0.163 0.15 6.301 5.98 9.09 21.395 Total 37070.58 100.00 0.593 0.36 6.416 100.00 100.00 100.000
156 Montcalm East
DMS Concentrates Product Separation SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg With Fines Concentrate +2.95 33308.70 74.50 2.117 0.82 4.393 97.56 93.11 67.417 Waste 2.95 11398.30 25.50 0.154 0.18 6.205 2.44 6.89 32.583 Total 44707.00 100.00 1.617 0.66 4.855 100.00 100.00 100.000 Without Fines Concentrate +2.95 30303.70 72.67 2.144 0.84 4.383 97.36 92.62 65.253 Waste 2.95 11398.30 27.33 0.154 0.18 6.205 2.64 7.38 34.747 Total 41702.00 100.00 1.600 0.66 4.881 100.00 100.00 100.000
Calculations
Crushed Material SG Size Fraction Weight Grade (%) Distribution (%) (mm) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 75+25 13241.60 78.97 2.685 0.78 3.781 72.769 81.84 87.142 25+6.7 3526.30 21.03 3.773 0.65 2.095 27.231 18.16 12.858 +3.1 total 16767.90 100.00 2.914 0.75 3.426 100.000 100.00 100.000 3.1+2.95 75+25 2538.00 58.82 0.612 0.56 6.141 43.576 38.65 62.020 25+6.7 1776.70 41.18 1.132 1.27 5.372 56.424 61.35 37.980 -3.1+2.95 total 4314.70 100.00 0.826 0.85 5.824 100.000 100.00 100.000 2.95+2.8 75+25 5914.20 69.96 0.107 0.10 6.424 53.002 48.23 70.667 25+6.7 2539.10 30.04 0.221 0.25 6.211 46.998 51.77 29.333 -2.95+2.8 total 8453.30 100.00 0.141 0.15 6.360 100.000 100.00 100.000 2.8 75+25 525.00 39.55 0.071 0.09 5.388 23.529 26.90 35.894 25+6.7 802.30 60.45 0.151 0.16 6.297 76.471 73.10 64.106 -2.8 total 1327.30 100.00 0.119 0.13 5.937 100.000 100.00 100.000 Fines 2705.00 100.00 1.825 0.64 4.501 100.000 100.00 100.000
Crushed Material - Summary SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 16767.90 49.95 2.914 0.75 3.426 83.217 64.95 36.735 3.1+2.95 4314.70 12.85 0.826 0.85 5.824 6.071 18.93 16.068 2.95+2.8 8453.30 25.18 0.141 0.15 6.360 2.034 6.31 34.375 2.8 1327.30 3.95 0.119 0.13 5.937 0.270 0.90 5.039 Fines 2705.00 8.06 1.825 0.64 4.501 8.408 8.91 7.784 Total 33568.20 100.00 1.749 0.58 4.659 100.000 100.00 100.000
Uncrushed Material SG Size Fraction Weight Grade (%) Distribution (%) (mm) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 75+25 4841.80 89.02 1.658 1.54 4.373 82.502 94.41 91.696 25+6.7 597.00 10.98 2.852 0.74 3.212 17.498 5.59 8.304 +3.1 total 5438.80 100.00 1.789 1.45 4.246 100.000 100.00 100.000 3.1+2.95 75+25 3585.00 94.78 0.699 0.31 7.274 88.769 85.70 96.092 25+6.7 197.30 5.22 1.607 0.94 5.376 11.231 14.30 3.908 -3.1+2.95 total 3782.30 100.00 0.746 0.34 7.175 100.000 100.00 100.000 2.95+2.8 75+25 963.20 75.42 0.274 0.33 5.638 72.557 55.86 73.922 25+6.7 313.90 24.58 0.318 0.80 6.103 27.443 44.14 26.078 -2.95+2.8 total 1277.10 100.00 0.285 0.45 5.752 100.000 100.00 100.000 2.8 75+25 204.70 60.10 0.104 0.18 5.192 48.104 61.46 61.301 25+6.7 135.90 39.90 0.169 0.17 4.937 51.896 38.54 38.699 -2.8 total 340.60 100.00 0.130 0.18 5.090 100.000 100.00 100.000 Fines 300.00 100.00 2.043 0.72 4.482 100.000 100.00 100.000
157 Uncrushed Material - Summary SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 5438.80 48.83 1.789 1.45 4.246 71.683 78.67 38.070 3.1+2.95 3782.30 33.96 0.746 0.34 7.175 20.797 12.92 44.743 2.95+2.8 1277.10 11.47 0.285 0.45 5.752 2.680 5.67 12.112 2.8 340.60 3.06 0.130 0.18 5.090 0.326 0.60 2.858 Fines 300.00 2.69 2.043 0.72 4.482 4.515 2.15 2.217 Total 11138.80 100.00 1.219 0.90 5.445 100.000 100.00 100.000
Crushed and Uncrushed - Combined SG Product Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 crushed 16767.90 75.51 2.914 0.75 3.426 83.39 61.51 71.332 uncrushed 5438.80 24.49 1.789 1.45 4.246 16.61 38.49 28.668 +3.1 total 22206.70 100.00 2.638 0.92 3.627 100.00 100.00 100.000 3.1+2.95 crushed 4314.70 53.29 0.826 0.85 5.824 55.80 73.93 48.079 uncrushed 3782.30 46.71 0.746 0.34 7.175 44.20 26.07 51.921 -3.1+2.95 total 8097.00 100.00 0.789 0.61 6.455 100.00 100.00 100.000 2.95+2.8 crushed 8453.30 86.88 0.141 0.15 6.360 76.65 68.31 87.979 uncrushed 1277.10 13.12 0.285 0.45 5.752 23.35 31.69 12.021 -2.95+2.8 total 9730.40 100.00 0.160 0.18 6.280 100.00 100.00 100.000 2.8 crushed 1327.30 79.58 0.119 0.13 5.937 78.16 74.55 81.967 uncrushed 340.60 20.42 0.130 0.18 5.090 21.84 25.45 18.033 -2.8 total 1667.90 100.00 0.122 0.14 5.764 100.00 100.00 100.000 Fines crushed 2705.00 90.02 1.825 0.64 4.501 88.96 88.91 90.055 uncrushed 300.00 9.98 2.043 0.72 4.482 11.04 11.09 9.945 Fines total 3005.00 100.00 1.847 0.65 4.499 100.00 100.00 100.000
With Fines SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 22206.70 66.67 2.638 0.92 3.627 83.07 74.78 55.042 3.1+2.95 8097.00 24.31 0.789 0.61 6.455 9.06 18.13 35.719 Fines 3005.00 9.02 1.847 0.65 4.499 7.87 7.10 9.239 +2.95 with fines 33308.70 100.00 2.117 0.82 4.393 100.00 100.00 100.000 2.95+2.8 9730.40 85.37 0.160 0.18 6.280 88.49 88.40 86.406 2.8 1667.90 14.63 0.122 0.14 5.764 11.51 11.60 13.594 Waste 11398.30 100.00 0.154 0.18 6.205 100.00 100.00 100.000
Without Fines SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 22206.70 73.28 2.638 0.92 3.627 90.17 80.49 60.645 3.1+2.95 8097.00 26.72 0.789 0.61 6.455 9.83 19.51 39.355 +2.95 total 30303.70 100.00 2.144 0.84 4.383 100.00 100.00 100.000 2.95+2.8 9730.40 85.37 0.160 0.18 6.280 88.49 88.40 86.406 2.8 1667.90 14.63 0.122 0.14 5.764 11.51 11.60 13.594 Waste 11398.30 100.00 0.154 0.18 6.205 100.00 100.00 100.000
DMS Concentrates Product Separation SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg With Fines Concentrate +2.95 33308.70 74.50 2.117 0.82 4.393 97.56 93.11 67.417 Waste 2.95 11398.30 25.50 0.154 0.18 6.205 2.44 6.89 32.583 Total 44707.00 100.00 1.617 0.66 4.855 100.00 100.00 100.000 Without Fines Concentrate +2.95 30303.70 72.67 2.144 0.84 4.383 97.36 92.62 65.253 Waste 2.95 11398.30 27.33 0.154 0.18 6.205 2.64 7.38 34.747 Total 41702.00 100.00 1.600 0.66 4.881 100.00 100.00 100.000
158 Montcalm West
DMS Concentrates Product Separation SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg With Fines Concentrate +2.95 10945.60 28.73 1.173 0.45 5.500 79.51 69.47 28.725 Waste 2.95 27148.30 71.27 0.122 0.08 5.502 20.49 30.53 71.275 Total 38093.90 100.00 0.424 0.19 5.501 100.00 100.00 100.000 Without Fines Concentrate +2.95 6780.60 19.98 1.585 0.59 5.357 76.46 65.00 19.563 Waste 2.95 27148.30 80.02 0.122 0.08 5.502 23.54 35.00 80.437 Total 33928.90 100.00 0.414 0.18 5.473 100.00 100.00 100.000
Calculations
Crushed Material SG Size Fraction Weight Grade (%) Distribution (%) (mm) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 75+25 1816.10 81.38 1.403 0.56 6.014 81.345 84.18 81.239 25+6.7 415.60 18.62 1.406 0.46 6.069 18.655 15.82 18.761 +3.1 total 2231.70 100.00 1.404 0.54 6.024 100.000 100.00 100.000 3.1+2.95 75+25 1505.10 62.84 0.907 0.23 5.689 57.648 49.94 61.408 25+6.7 889.90 37.16 1.127 0.39 6.047 42.352 50.06 38.592 -3.1+2.95 total 2395.00 100.00 0.989 0.29 5.822 100.000 100.00 100.000 2.95+2.8 75+25 10587.40 68.54 0.145 0.11 5.857 64.884 77.40 65.505 25+6.7 4858.70 31.46 0.171 0.07 6.721 35.116 22.60 34.495 -2.95+2.8 total 15446.10 100.00 0.153 0.10 6.129 100.000 100.00 100.000 2.8 75+25 4733.10 50.86 0.036 0.04 2.895 45.850 57.98 38.243 25+6.7 4573.60 49.14 0.044 0.03 4.838 54.150 42.02 61.757 -2.8 total 9306.70 100.00 0.040 0.04 3.850 100.000 100.00 100.000 Fines 3435.00 100.00 0.379 0.17 5.803 100.000 100.00 100.000
Crushed Material - Summary SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 2231.70 6.80 1.404 0.54 6.024 32.834 27.99 7.561 3.1+2.95 2395.00 7.30 0.989 0.29 5.822 24.823 16.06 7.842 2.95+2.8 15446.10 47.07 0.153 0.10 6.129 24.801 34.86 53.238 2.8 9306.70 28.36 0.040 0.04 3.850 3.896 7.56 20.150 Fines 3435.00 10.47 0.379 0.17 5.803 13.647 13.53 11.210 Total 32814.50 100.00 0.291 0.13 5.419 100.000 100.00 100.000
Uncrushed Material SG Size Fraction Weight Grade (%) Distribution (%) (mm) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 75+25 1330.60 92.33 3.274 0.85 3.506 91.897 93.25 95.214 25+6.7 110.60 7.67 3.473 0.74 2.120 8.103 6.75 4.786 +3.1 total 1441.20 100.00 3.289 0.84 3.400 100.000 100.00 100.000 3.1+2.95 75+25 651.30 91.38 0.659 1.35 5.742 84.482 96.89 92.581 25+6.7 61.40 8.62 1.284 0.46 4.881 15.518 3.11 7.419 -3.1+2.95 total 712.70 100.00 0.713 1.27 5.668 100.000 100.00 100.000 2.95+2.8 75+25 1904.10 86.97 0.228 0.14 8.260 78.897 82.37 88.714 25+6.7 285.30 13.03 0.407 0.20 7.013 21.103 17.63 11.286 -2.95+2.8 total 2189.40 100.00 0.251 0.15 8.098 100.000 100.00 100.000 2.8 75+25 0.00 0.00 0.000 0.00 0.000 0.000 0.00 0.000 25+6.7 206.10 100.00 0.101 0.05 5.526 100.000 100.00 100.000 -2.8 total 206.10 100.00 0.101 0.05 5.526 100.000 100.00 100.000 Fines 730.00 100.00 1.076 0.44 5.391 100.000 100.00 100.000
159 Uncrushed Material - Summary SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 1441.20 27.30 3.289 0.84 3.400 71.770 43.70 15.436 3.1+2.95 712.70 13.50 0.713 1.27 5.668 7.692 32.70 12.726 2.95+2.8 2189.40 41.47 0.251 0.15 8.098 8.331 11.66 55.852 2.8 206.10 3.90 0.101 0.05 5.526 0.315 0.37 3.588 Fines 730.00 13.83 1.076 0.44 5.391 11.892 11.57 12.398 Total 5279.40 100.00 1.251 0.53 6.012 100.000 100.00 100.000
Crushed and Uncrushed - Combined SG Product Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 crushed 2231.70 60.76 1.404 0.54 6.024 39.79 49.90 73.290 uncrushed 1441.20 39.24 3.289 0.84 3.400 60.21 50.10 26.710 +3.1 total 3672.90 100.00 2.143 0.66 4.994 100.00 100.00 100.000 3.1+2.95 crushed 2395.00 77.07 0.989 0.29 5.822 82.34 43.31 77.538 uncrushed 712.70 22.93 0.713 1.27 5.668 17.66 56.69 22.462 -3.1+2.95 total 3107.70 100.00 0.925 0.52 5.787 100.00 100.00 100.000 2.95+2.8 crushed 15446.10 87.59 0.153 0.10 6.129 81.13 82.30 84.226 uncrushed 2189.40 12.41 0.251 0.15 8.098 18.87 17.70 15.774 -2.95+2.8 total 17635.50 100.00 0.165 0.10 6.373 100.00 100.00 100.000 2.8 crushed 9306.70 97.83 0.040 0.04 3.850 94.70 96.94 96.919 uncrushed 206.10 2.17 0.101 0.05 5.526 5.30 3.06 3.081 -2.8 total 9512.80 100.00 0.041 0.04 3.886 100.00 100.00 100.000 Fines crushed 3435.00 82.47 0.379 0.17 5.803 62.37 64.51 83.512 uncrushed 730.00 17.53 1.076 0.44 5.391 37.63 35.49 16.488 Fines total 4165.00 100.00 0.501 0.22 5.731 100.00 100.00 100.000
With Fines SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 3672.90 33.56 2.143 0.66 4.994 61.33 49.14 30.474 3.1+2.95 3107.70 28.39 0.925 0.52 5.787 22.41 32.49 29.875 Fines 4165.00 38.05 0.501 0.22 5.731 16.26 18.37 39.652 +2.95 with fines 10945.60 100.00 1.173 0.45 5.500 100.00 100.00 100.000 2.95+2.8 17635.50 64.96 0.165 0.10 6.373 88.14 84.44 75.249 2.8 9512.80 35.04 0.041 0.04 3.886 11.86 15.56 24.751 Waste 27148.30 100.00 0.122 0.08 5.502 100.00 100.00 100.000
Without Fines SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg +3.1 3672.90 54.17 2.143 0.66 4.994 73.24 60.20 50.496 3.1+2.95 3107.70 45.83 0.925 0.52 5.787 26.76 39.80 49.504 +2.95 total 6780.60 100.00 1.585 0.59 5.357 100.00 100.00 100.000 2.95+2.8 17635.50 64.96 0.165 0.10 6.373 88.14 84.44 75.249 2.8 9512.80 35.04 0.041 0.04 3.886 11.86 15.56 24.751 Waste 27148.30 100.00 0.122 0.08 5.502 100.00 100.00 100.000
DMS Concentrates Product Separation SG Weight Grade (%) Distribution (%) (g) (%) Ni Cu Mg Ni Cu Mg With Fines Concentrate +2.95 10945.60 28.73 1.173 0.45 5.500 79.51 69.47 28.725 Waste 2.95 27148.30 71.27 0.122 0.08 5.502 20.49 30.53 71.275 Total 38093.90 100.00 0.424 0.19 5.501 100.00 100.00 100.000 Without Fines Concentrate +2.95 6780.60 19.98 1.585 0.59 5.357 76.46 65.00 19.563 Waste 2.95 27148.30 80.02 0.122 0.08 5.502 23.54 35.00 80.437 Total 33928.90 100.00 0.414 0.18 5.473 100.00 100.00 100.000
160 Appendix 3: Assay Values
161 Craig 8112 t t /mt /mt Pt Pt Pt Pt 0.111 0.109 3 0.124 8 0.085 Assay Assay Assay Assay % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g % g/mt g/mt g/mt g Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd 2.8 353.3 0.005 0.08 1.99 0.112 0 0.5 0.03 0.03 2.8 1299.7 0.011 0.1 6.592 0.212 0.02 0.3 0.07 0.12 2.8 488.8 0.01 0.09 1.651 0.308 0 0.5 0.06 0.08 2.8 199.4 0.008 0.18 3.218 0.187 0.03 1.4 0.04 0.05 +3.1 5591.4 0.048 1.05 4.803 1.717 0.35 2.8 0.21 0.16 +3.1 1453.6 0.062 1.4 4.134 2.343 0.04 3.8 0.16 0.13 +3.1 7791.4 0.061 0.89 3.754 2.275 0.23 2.4 0.14 0.14 +3.1 1252.3 0.095 0.81 1.93 3.53 0.02 2.7 0.23 0.2 75+25 (gm)Co Cu Mg 75+25 (gm)Co Cu Mg 25+6.7 (gm)Co Cu Mg 25+6.7 (gm)Co Cu Mg Crushed % % % Crushed % % % 3.1+2.95 2.95+2.8 7675 569.4 0.011 0.004 0.15 0.05 7.89 3.499 0.241 0.073 0.03 0.01 0.6 0.2 0.09 0.01 0.11 0.02 3.1+2.95 2.95+2.8 2300.3 216.2 0.013 0.005 0.16 0.08 6.58 2.405 0.276 0.103 0.02 0.01 0.4 0.4 0.08 0.05 0.1 0.04 3.1+2.95 2.95+2.8 5286.7 1569 0.011 0.011 0.11 0.17 7.899 7.116 0.154 0.208 0.01 0.03 0.4 1.5 0.03 0.03 0.02 0.03 3.1+2.95 2.95+2.8 561.8 852.4 0.023 0.011 0.43 0.14 5.475 6.384 0.704 0.198 0.02 0 1.2 0.5 0.07 0.02 0.06 0.02 Uncrushed % % % Uncrushed % % % Head Head Grade 14189.1 0.025 0.499 6.350 0.813 0.155 1.448 0.13 Head Head Grade 5269.8 0.026 0.484 5.737 0.823 0.025 1.313 0.098 Head Head Grade 15135.9 0.037 0.517 5.482 1.256 0.125 1.547 0.08 Head Head Grade 2865.9 0.050 0.492 4.039 1.752 0.015 1.661 0.123 DMS DMS TestResults Size (mm)Size Weight Sinks Size (mm)Size weight Size (mm)Size weight Size (mm)Size weight
162 t t /mt /mt Pt Pt Pt Pt 0.120 0.095 4 0.103 0 0.092 Assay Assay Assay Assay % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g % g/mt g/mt g/mt g Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd Tails 1168.1 0.004 0.1 3.631 0.068 0.04 0.5 0.14 0.09 Tails 3363.3 0.014 0.33 6.418 0.342 0.05 0.8 0.07 0.09 Tails 347 0.005 0.21 3.31 0.126 0.03 0.7 0.02 0.03 Tails 1519.8 0.021 0.29 6.317 0.642 0.02 0.9 0.07 0.05 Conc 9975.6 0.043 0.75 4.573 1.512 0.03 1.9 0.14 0.12 Conc 1240.6 0.06 1.1 4.106 2.273 0.06 3.1 0.13 0.2 Conc 11785.5 0.036 0.32 6.084 1.209 0.08 1.1 0.12 0.1 Conc 1050.9 0.075 0.98 2.988 2.839 0.04 3.2 0.2 0.16 Tails 2 2876 0.008 0.1 7.905 0.139 0.02 0.4 0.06 0.05 Tails 2 Tails 2 923.4 0.009 0.2 7.996 0.152 0.02 0.5 0.01 0.01 Tails 2 75+25 (gm) Co Cu Mg 75+25 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg Crushed % % % Crushed % % % Uncrushed % % % Uncrushed % % % Head Head Grade 14019.7 0.033 0.563 5.178 1.110 0.029 1.476 0.12 Head Head Grade 4603.9 0.026 0.537 5.795 0.862 0.053 1.420 0.086 Head Head Grade 13055.9 0.033 0.309 6.146 1.105 0.074 1.047 0.11 Head Head Grade 2570.7 0.043 0.572 4.956 1.540 0.028 1.840 0.123 Sorting Test Results Size (mm)Size Weight Size (mm)Size weight Size (mm)Size weight Size (mm)Size weight
163 6 5 093 0.125 g/mt g/mt g/mt g/mt g/mt g/mt g/mt g/mt i Au Ag Pd Pt i Au Ag Pd Pt Assay Assay % % % % % % % % (gm) Co Cu Mg N (gm) Co Cu Mg N weight weight Total 77164.5 0.033 0.480 4.823 1.113 0.079 1.419 0.104 0.09 Crush 2553.6 0.036Crush 0.68 5.782 40636.1 1.28 0.029 0.08 0.508 4.280 2.5 0.908 0.12 0.075 0.12 1.392 0.104 0.09 Uncrush 2900 0.061 0.57 3.685 2.253 0.04 2.2 0.12 0.13 Uncrush 36528.4 0.039 0.449 5.426 1.340 0.085 1.450 0.103 0. Fines 6.7 Head Head Grade 5453.6 0.049 0.622 4.667 1.797 0.059 2.340 0.120 Head Grades
164 Craig LGBX t t /mt /mt Pt Pt Pt Pt .137 .089 4 0.177 0 0.093 Assay Assay Assay Assay % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g % g/mt g/mt g/mt g Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd 2.8 105 0.01 0.27 2.533 0.291 0.01 0.7 0.04 0.03 2.8 1060 0.006 0.09 1.075 0.184 0 0.2 0.02 0.02 2.8 2546.1 0.005 0.15 1.232 0.159 0.01 0.2 0.04 0.03 2.8 4488 0.006 0.15 1.007 0.181 0.03 0.3 0.11 0.11 +3.1 266.1 0.128 0.57 1.24 5.226 0.01 1.7 0.2 0.22 +3.1 6394.9 0.16 0.61 0.383 5.523 0 1 0.19 0.3 +3.1 4432 0.129 0.31 1.529 5.182 0.01 0.9 0.07 0.18 +3.1 11263.1 0.106 0.4 2.239 4.319 0.03 1.3 0.18 0.13 75+25 (gm) Co Cu Mg 75+25 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg Crushed % % % Crushed % % % 3.1+2.95 2.95+2.8 141.7 185.1 0.031 0.012 0.76 0.3 3.804 3.379 1.184 0.32 0.02 0.04 1.5 0.8 0.18 0.06 0.15 0.07 3.1+2.95 2.95+2.8 2967.7 1416 0.009 0.012 0.23 0.28 3.856 3.078 0.164 0.281 0.01 0.02 0.4 0.5 0.03 0.1 0.02 0.07 3.1+2.95 2.95+2.8 994.7 3039.2 0.014 0.01 0.19 0.28 5.374 4.322 0.35 0.208 0.01 0.02 0.4 0.6 0.04 0.04 0.02 0.03 3.1+2.95 2.95+2.8 4742.7 3520.1 0.011 0.01 0.14 0.13 4.909 3.309 0.191 0.284 0.02 0.02 0.2 0.2 0.03 0.06 0.02 0.05 Uncrushed % % % Uncrushed % % % Head GradeHead 697.9 0.060 0.492 2.522 2.362 0.020 1.270 0.135 0 Head GradeHead 11838.6 0.091 0.429 1.638 3.075 0.005 0.718 0.12 Head GradeHead 11012 0.057 0.254 2.578 2.211 0.013 0.610 0.052 0 Head GradeHead 24013.9 0.054 0.262 2.693 2.139 0.027 0.735 0.12 DMS Test Results Size (mm) Size weight Size (mm) Size weight Size (mm) Size weight Size (mm) Size WeightSinks
165 t t /mt /mt Pt Pt Pt Pt .131 0.128 0.073 2 0.084 Assay Assay Assay Assay % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g % g/mt g/mt g/mt g Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd Tails 471.9 0.03 0.75 2.903 1.064 0.04 1.8 0.1 0.1 Tails 210.2 0.007 1.34 3.107 0.153 0.59 4.5 0.12 0.07 Tails 403.5 0.003 0.05 1.354 0.071 0.01 0.1 0.03 0.02 Tails 4370.2 0.02 0.21 3.21 0.669 0.01 0.4 0.05 0.04 Conc 377.4 0.078 1.59 1.818 2.809 0.02 3.2 0.23 0.17 Conc 8348.3 0.039 0.42 2.201 1.396 0.01 0.9 0.13 0.13 Conc 20720 0.06 0.33 2.617 2.481 0.01 0.8 0.11 0.09 Conc 5168.9 0.082 0.32 2.117 3.25 0.01 1 0.1 0.1 Tails2 Tails2 18.8 Tails2 2311.4 0.009 0.25 4.495 0.22 0.02 0.5 0.04 0.04 Tails2 75+25 (gm) Co Cu Mg 75+25 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg Crushed % % % Crushed % % % Uncrushed % % % Uncrushed % % % Head GradeHead 849.3 0.051 1.123 2.421 1.839 0.031 2.422 0.158 0 Head GradeHead 8577.3 0.038 0.442 2.218 1.362 0.024 0.986 0.129 Head GradeHead 23434.9 0.054 0.317 2.780 2.217 0.011 0.758 0.10 Head GradeHead 9539.1 0.054 0.270 2.618 2.068 0.010 0.725 0.077 Sorting Test Results Size (mm) Size weight Size (mm) Size Weight Size (mm) Size weight Size (mm) Size weight
166 0.137 0.184 1.441 0.082 g/mt g/mt g/mt g/mt iAu Ag Pd Pt Assay 3.381 1.923 0.422 0.083 % % % % (gm) Co Cu Mg N 6737 weight Crush 5785 0.078 0.34 2.064 3.2 0.09 1.3 0.15 0.11 Uncrush 952 0.11 0.92 1.067 4.48 0.03 2.3 0.39 0.3 Fines Fines 6.7 Head Grade Head
167 Fraser Nickel t t /mt /mt Pt Pt Pt Pt 5 0.091 9 0.090 79 0.108 08 0.159 Assay Assay Assay Assay % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g % g/mt g/mt g/mt g Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd 2.7 35.01 0.006 0.08 2.895 0.087 0.09 0.4 0.02 0 2.7 12.54 0.005 0.04 3.083 0.032 0.02 0.4 0 0 2.7 19.82 0.006 0.17 2.617 0.08 0.05 0.8 0.01 0 2.7 249.43 0.006 0.05 2.816 0.105 0.04 0.3 0.04 0.03 +3.1 327.58 0.047 1.03 3.184 1.513 0.06 2.8 0.17 0.25 +3.1 755.5 0.051 2.8 1.625 1.979 0.15 7.7 0.24 0.4 +3.1 1520.6 0.05 0.72 1.999 1.787 0.04 2.1 0.17 0.23 +3.1 1621.6 0.064 0.86 1.658 2.318 0.05 2.6 0.15 0.32 75+25 (gm) Co Cu Mg 75+25 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg 3.1+2.9 2.9+2.7 380.14 1213.8 0.02 0.009 0.35 0.11 5.836 4.369 0.604 0.152 0.03 0.02 1 0.3 0.01 0.05 0.02 0.19 3.1+2.9 2.9+2.7 6870 2699.2 0.021 0.01 0.37 0.15 3.927 4.26 0.706 0.2 0.14 0.01 0.9 0.5 0.08 0.03 0.11 0.02 3.1+2.9 2.9+2.7 2746.4 6975.2 0.03 0.015 0.79 0.17 3.52 5.971 1.081 0.317 0.05 0.06 2.3 0.6 0.07 0.11 0.27 0.1 3.1+2.9 2.9+2.7 1381 6434.7 0.027 0.01 0.45 0.11 5.006 4.614 0.888 0.201 0.02 0.02 1.2 0.4 0.07 0.02 0.11 0.03 Crushed % % % Crushed % % % Uncrushed % % % Uncrushed % % % Head GradeHead 1956.53 0.017 0.310 4.429 0.467 0.030 0.856 0.04 Head GradeHead 10337.24 0.020 0.490 3.845 0.666 0.107 1.292 0.0 Head GradeHead 11262.02 0.023 0.395 4.831 0.701 0.055 1.217 0.1 Head GradeHead 9686.73 0.021 0.282 4.129 0.651 0.026 0.880 0.04 DMS Test Results Size (mm) Size WeightSinks Size (mm) Size weight Size (mm) Size weight Size (mm) Size weight
168 t t /mt /mt Pt Pt Pt Pt 0.091 0.084 2 0.238 60 0.070 Assay Assay Assay Assay % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g % g/mt g/mt g/mt g Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd Tails 1041.3 0.013 0.18 4.737 0.257 0 0.8 0.02 0.03 Tails 1438.5 0.008 0.03 10.248 0.05 0.04 0.4 0 0 Tails 12.55 0.009 0.11 6.499 0.15 0.02 0.6 0 0.03 Tails 4800 0.016 0.23 5.058 0.368 0.04 0.7 0.03 0.06 Conc 304.6 0.028 1.34 3.535 1.084 0.02 8.8 0.27 0.34 Conc 8747.4 0.034 0.46 3.62 1.056 0.05 1.9 0.12 0.28 Conc 14200 0.026 0.32 3.929 0.861 0.02 1.1 0.06 0.07 Conc 2722.9 0.036 0.38 3.014 1.063 0.04 1.3 0.08 0.11 Tails2 237 0.018 0.18 4.366 0.483 0 0.7 0.03 0.04 Tails2 123.2 0.008 0.19 9.106 0.093 0.02 0.6 0.04 0.06 Tails2 46.07 0.025 0.45 4.552 0.793 0.03 1.1 0.08 0.06 Tails2 1604.2 0.026 0.48 3.504 0.812 0.04 1.3 0.08 0.11 75+25 (gm) Co Cu Mg 75+25 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg Crushed % % % Crushed % % % Uncrushed % % % Uncrushed % % % Head GradeHead 1582.9 0.017 0.403 4.450 0.450 0.004 2.324 0.070 Head GradeHead 10309.1 0.030 0.397 4.610 0.904 0.048 1.675 0.10 Head GradeHead 14258.62 0.026 0.320 3.933 0.860 0.020 1.100 0.0 Head GradeHead 9127.1 0.024 0.319 4.175 0.653 0.040 0.984 0.054 Sorting Test Results Size (mm) Size weight Size (mm) Size Weight Size (mm) Size weight Size (mm) Size weight
169 0.105 0.099 1.312 0.054 g/mt g/mt g/mt g/mt iAu Ag Pd Pt Assay 0.841 3.771 0.499 0.027 % % % % (gm) Co Cu Mg N weight 6671.93 Crush 4725.94 0.025 0.56 3.846 0.782 0.06 1.4 0.07 0.09 Uncrush 1945.99 0.031 0.35 3.59 0.984 0.04 1.1 0.17 0.14 Fines Fines 6.7 Head Grade Head
170 Fraser Copper t t /mt /mt Pt Pt Pt Pt 6 0.856 34 2.740 27 0.464 801 1.977 Assay Assay Assay Assay % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g % g/mt g/mt g/mt g Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd 2.7 2.7 1041.8 0.002 1.74 1.153 0.03 0.03 4.1 0.16 0.09 2.7 175.82 0.003 0.06 1.858 0.008 0.01 1.9 0 0 2.7 2421 0.003 0.17 1.731 0.01 0.03 1.5 0 0.01 +3.1 2642.2 0.015 26.8 0.061 1.664 0.21 89 4.52 3.85 +3.1 2091.6 0.006 23.1 0.444 0.412 0.13 46 3.23 2.24 +3.1 3806.8 0.007 25.81 0.202 0.497 0.17 59 5.75 4.05 +3.1 2162.5 0.006 28.36 0.09 0.417 0.16 70 5.69 3.83 75+25 (gm) Co Cu Mg 75+25 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg 3.1+2.9 2.9+2.7 113.4 986.6 0.009 0.004 5.34 0.35 2.329 2.548 0.542 0.037 2.73 0.01 50 5.1 0.97 0.05 0.55 0.02 3.1+2.9 2.9+2.7 202.9 8283.9 0.007 0.004 1.66 0.34 4.459 2.954 0.079 0.034 0.06 0.04 4.9 3.1 0.16 0.04 0.19 0.07 3.1+2.9 2.9+2.7 403.8 3497.8 0.005 0.003 0.58 0.34 2.182 2.704 0.246 0.04 0.04 0.16 2.6 6.3 0.13 0.04 0.16 0.03 3.1+2.9 2.9+2.7 436.9 4870 0.007 0.004 0.49 0.41 5.224 3.69 0.076 0.027 0.06 0.04 4.5 8.1 0.06 0.06 0.04 0.03 Crushed % % % Crushed % % % Size (mm) Size WeightSinks Size (mm) Size weight Size (mm) Size weight Size (mm) Size weight Uncrushed % % % Uncrushed % % % Head GradeHead 3742.2 0.012 19.176 0.785 1.201 0.234 65.699 3.2 Head GradeHead 11620.2 0.004 4.585 2.367 0.102 0.056 10.943 0.6 Head GradeHead 7884.22 0.005 12.644 1.450 0.270 0.155 31.458 2. Head GradeHead 9890.4 0.004 6.466 2.491 0.110 0.065 19.860 1.27
DMS Test Results
171 t t /mt /mt Pt Pt Pt Pt 1 0.800 5 0.322 46 2.449 57 2.160 Assay Assay Assay Assay % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g % g/mt g/mt g/mt g Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd Tails 3200 0.011 16.44 0.975 0.984 0.13 80 2.36 1.98 Tails 2899.4 0.004 1.19 2.324 0.149 0.23 30 0.59 1.41 Tails 6200 0.005 4.99 2.777 0.103 0.03 14 0.58 0.6 Tails 6870.1 0.004 0.2 3.425 0.071 0.04 2.5 0.05 0.1 Conc 1040.3 0.011 26.54 0.153 1.155 0.11 125 5.14 4.07 Conc 5052.6 0.026 25.4 0.111 3.285 0.35 115 3.59 2.96 Conc 638.5 0.006 22.72 1.375 0.341 0.08 57 2.65 2.19 Conc 2127.5 0.004 12.52 1.272 0.179 0.05 25 1.21 1.19 Tails2 614.3 0.011 17.09 1.283 1.071 0.09 75 2.28 2.15 Tails2 987.2 0.005 2.06 2.262 0.438 0.26 8.5 0.33 0.27 Tails2 545.3 0.004 13.99 1.566 0.187 0.05 30 1.28 1.44 Tails2 1227.8 0.004 0.69 3.259 0.068 0.02 4.7 0.08 0.06 75+25 (gm) Co Cu Mg 75+25 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg Crushed % % % Crushed % % % Size (mm) Size weight Size (mm) Size weight Size (mm) Size weight Size (mm) Size Weight Uncrushed % % % Uncrushed % % % Head GradeHead 4854.6 0.011 18.687 0.838 1.032 0.121 89.010 2.9 Head GradeHead 8939.2 0.017 14.970 1.066 1.953 0.301 75.669 2.2 Head GradeHead 7383.8 0.005 7.188 2.566 0.130 0.036 18.900 0.81 Head GradeHead 10225.4 0.004 2.822 2.957 0.093 0.040 7.446 0.29 Sorting TestResults
172 4.557 2.883 74.731 0.252 g/mt g/mt g/mt g/mt iAu Ag Pd Pt Assay 1.085 0.945 19.173 0.012 % % % % (gm) Co Cu Mg N weight 13464.39 Crush 5094.47 0.005 11.03 1.937 0.187 0.24 25 1.59 4.19 Uncrush 8369.92 0.016 24.13 0.341 1.632 0.26 105 3.67 4.78 Fines Fines 6.7 Head Grade Head
173 Thayer Lindsley Footwall t t /mt /mt Pt Pt Pt Pt 0.927 2 1.119 17 1.025 819 1.492 Assay Assay Assay Assay % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g % g/mt g/mt g/mt g Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd 2.7 272.3 0.002 0.07 0.743 0.019 0.06 0.3 0.02 0.01 2.7 1600 0.001 0.08 0.37 0.009 0.29 0.7 0.02 0.02 2.7 1400 0.002 0.12 0.868 0.014 0.19 0.3 0.01 0.03 2.7 890.1 0.001 0.04 0.12 0.034 0.05 0.1 0.01 0 +3.1 1600 0.075 16.74 0.246 1.873 0.29 42 3.88 1.4 +3.1 1200 0.091 13.39 0.268 2.272 0.26 29 3.97 2.29 +3.35 5200 0.124 6.19 0.388 3.094 0.23 14.4 3.15 1.46 +3.35 8200 0.07 16.67 0.239 1.711 0.74 31 4.47 2.33 75+25 (gm) Co Cu Mg 75+25 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg 3.1+2.9 2.9+2.7 126.3 227.1 0.011 0.004 1.6 0.91 3.221 1.448 0.161 0.061 0.05 0.95 4.4 1.7 0.26 0.1 0.11 0.11 3.1+2.9 2.9+2.7 389.7 2800 0.007 0.011 0.04 0.22 3.892 2.953 0.075 0.189 0.02 0.03 0.5 1 0.02 0.2 0 0.99 3.1+2.9 2.9+2.7 495.4 3000 0.016 0.006 1.62 0.56 4.37 3.606 0.254 0.064 0.32 0.43 7 1.5 0.51 0.09 0.33 0.03 3.1+2.9 2.9+2.7 886.5 2600 0.015 0.006 1.58 0.89 5.264 2.994 0.223 0.065 0.08 0 3.8 1.8 0.3 0.09 0.41 0.06 Crushed % % % Crushed % % % 3.35+3.1 969 0.034 4.04 2.835 0.798 0.09 14.9 1.18 4.64 3.35+3.1 685.2 0.027 2.13 3.647 0.656 0.07 5.5 0.32 0.23 Uncrushed % % % Uncrushed % % % Head GradeHead 2225.7 0.055 12.226 0.598 1.364 0.316 30.653 2.8 Head GradeHead 5989.7 0.024 2.809 1.786 0.551 0.145 6.497 0.896 Head GradeHead 11064.4 0.064 3.503 1.714 1.554 0.271 8.831 1.63 Head GradeHead 13261.8 0.047 10.700 1.283 1.122 0.470 20.066 2. DMS Test Results Size (mm) Size weight Size (mm) Size weight Size (mm) Size weight Size (mm) Size WeightSinks
174 t t /mt /mt Pt Pt Pt Pt .887 1.403 1.113 5 0.801 Assay Assay Assay Assay % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g % g/mt g/mt g/mt g Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd Tails 1000 0.019 15.68 0.801 0.417 0.36 31 2 0.68 Tails 6400 0.006 0.66 4.627 0.043 0.02 1.1 0.03 0.25 Tails 4000 0.023 4.03 2.557 0.45 0.88 9 0.53 0.41 Tails 3000 0.008 3.61 3.316 0.15 0.49 8.6 0.55 0.21 Conc 1400 0.084 12.16 0.363 2.096 2.26 28 3.8 1.92 Conc 7200 0.088 14.26 0.215 2.135 0.23 33 5.19 1.88 Conc 4400 0.099 5.6 0.972 2.435 0.17 13.4 2.18 1.32 Conc 12800 0.063 13.75 1.562 1.566 0.35 29 2.42 0.94 Tails2 Tails2 Tails2 Tails2 75+25 (gm) Co Cu Mg 75+25 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg Crushed % % % Crushed % % % Uncrushed % % % Uncrushed % % % Head GradeHead 2400 0.057 13.627 0.546 1.396 1.468 29.250 3.050 Head GradeHead 13600 0.049 7.860 2.291 1.151 0.131 17.988 2.762 Head GradeHead 8400 0.063 4.852 1.727 1.490 0.508 11.305 1.394 0 Head GradeHead 15800 0.053 11.825 1.895 1.297 0.377 25.127 2.06 Sorting Test Results Size (mm) Size weight Size (mm) Size weight Size (mm) Size weight Size (mm) Size Weight
175 1.375 2.655 16.750 0.305 g/mt g/mt g/mt g/mt iAu Ag Pd Pt Assay 1.275 1.686 8.130 0.052 % % % % (gm) Co Cu Mg N 4800 weight Crush 2400 0.056 6.23 2.287 1.353 0.39 14.2 2.79 1.91 Uncrush 2400 0.048 10.03 1.085 1.197 0.22 19.3 2.52 0.84 Fines Fines 6.7 Head Grade Head
176 Thayer Lindsley Zone 1
t t /mt /mt Pt Pt Pt Pt 6 0.072 0 0.042 82 0.092 73 0.033 Assay Assay Assay Assay % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g % g/mt g/mt g/mt g Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd 2.7 43.07 0.004 0.3 1.525 0.065 0.4 2 0.06 0 2.7 0 2.7 14.88 0.008 0.18 5.054 0.125 0.36 0.8 0.01 0.07 2.7 0 +3.1 1145.3 0.079 1.1 3.827 2.034 0.06 2.8 0.29 0.15 +3.1 5890.8 0.049 0.56 4.982 1.268 0.11 1.6 0.13 0.16 +3.1 670.7 0.055 0.38 5.258 1.479 0.13 1.3 0.29 0.04 +3.1 6900.16 0.032 0.59 6.439 0.844 0.42 1.8 0.12 0.03 75+25 (gm) Co Cu Mg 75+25 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg 3.1+2.9 2.9+2.7 903.38 1513.61 0.019 0.012 0.45 0.18 6.496 6.303 0.351 0.207 0.16 0.1 1.6 0.7 0.11 0.06 0.03 0.04 3.1+2.9 2.9+2.7 6480.73 2139.5 0.015 0.01 0.25 0.22 7.273 5.952 0.263 0.169 0.09 0.23 0.7 0.5 0.05 0.05 0.04 0.06 3.1+2.9 2.9+2.7 2120.11 1679.24 0.018 0.009 0.21 0.15 8.208 6.883 0.349 0.131 0.06 0.15 0.6 0.9 0.06 0.05 0.02 0.07 3.1+2.9 2.9+2.7 4883.2 2685.9 0.015 0.009 0.13 0.07 7.141 6.298 0.266 0.155 0.09 0.15 0.5 0.2 0.03 0.03 0.05 0.01 Crushed % % % Crushed % % % Size (mm) Size weight Size (mm) Size weight Size (mm) Size weight Size (mm) Size WeightSinks Uncrushed % % % Uncrushed % % % Head GradeHead 3605.36 0.035 0.541 5.508 0.822 0.106 1.608 0.14 Head GradeHead 14511.03 0.028 0.371 6.148 0.657 0.119 1.036 0.0 Head GradeHead 4484.93 0.020 0.213 7.260 0.436 0.105 0.818 0.09 Head GradeHead 14469.26 0.022 0.338 6.650 0.521 0.259 1.064 0.0
DMS Test Results
177 t t /mt /mt Pt Pt Pt Pt 199 046 .100 .102 Assay Assay Assay Assay % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g % g/mt g/mt g/mt g Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd Tails 3200 0.02 0.54 5.71 0.412 0.09 1.6 0.14 0.24 Tails 6400 0.015 0.36 6.401 0.275 0.03 1.2 0.09 0.11 Tails 4800 0.016 0.25 7.18 0.303 0.11 0.7 0.1 0.04 Tails 8000 0.026 0.45 6.023 0.677 0.36 1.4 0.06 0.13 Conc 1200 0.071 0.41 4.056 1.725 0.16 1.5 0.29 0.09 Conc 6200 0.046 0.67 4.737 1.165 0.07 1.8 0.2 0.09 Conc 400 0.039 0.33 5.951 1.017 0.06 1.3 0.2 0.12 Conc 5200 0.009 0.1 7.795 0.127 0.31 0.4 0.05 0.06 Tails2 Tails2 Tails2 Tails2 75+25 (gm) Co Cu Mg 75+25 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg Crushed % % % Crushed % % % Size (mm) Size weight Size (mm) Size weight Size (mm) Size weight Size (mm) Size Weight Uncrushed % % % Uncrushed % % % Head GradeHead 4400 0.034 0.505 5.259 0.770 0.109 1.573 0.181 0. Head GradeHead 12600 0.030 0.513 5.582 0.713 0.050 1.495 0.144 0 Head GradeHead 5200 0.018 0.256 7.085 0.358 0.106 0.746 0.108 0. Head GradeHead 13200 0.019 0.312 6.721 0.460 0.340 1.006 0.056 0 Sorting TestResults
178 0.185 0.234 2.412 0.050 g/mt g/mt g/mt g/mt iAu Ag Pd Pt Assay 1.542 4.492 0.697 0.061 % % % % (gm) Co Cu Mg N 8600 weight Crush 2800 0.036 0.38 6.148 0.916 0.07 1.4 0.14 0.05 Uncrush 5800 0.073 0.85 3.693 1.844 0.04 2.9 0.28 0.25 Fines Fines 6.7 Head Grade Head
179 Thayer Lindsley Zone 2
t t /mt /mt Pt Pt Pt Pt 343 .496 0.333 8 0.540 Assay Assay Assay Assay % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g % g/mt g/mt g/mt g Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd 2.7 18.2 0.012 0.48 2.288 0.219 0.06 6 0.04 0.02 2.7 825 0.002 0.09 1.276 0.048 0.01 0.3 0.01 0.01 2.7 195 0.001 0.01 0.105 0.046 0.01 0.1 0 0 2.7 53.1 0.002 0.05 0.736 0.02 0.01 0.2 0.02 0 +3.1 9600 0.06 1.32 3.638 2.04 0.04 5.5 0.34 0.95 +3.1 4200 0.078 1.48 2.803 2.557 0.03 6.2 0.52 0.96 +3.1 5200 0.062 1.24 2.658 2.104 0.07 5.2 0.25 0.37 +3.1 757.2 0.086 1.27 2.36 2.66 0.05 5.7 0.69 0.59 75+25 (gm) Co Cu Mg 75+25 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg 3.1+2.9 2.9+2.7 3800 4200 0.009 0.007 0.06 0.1 5.642 4.334 0.134 0.11 0.02 0.02 0.3 0.5 0.04 0.06 0.06 0.04 3.1+2.9 2.9+2.7 1810 3800 0.025 0.008 0.75 0.23 5.219 4.923 0.739 0.15 0.04 0.04 3.1 1 0.19 0.04 0.39 0.14 3.1+2.9 2.9+2.7 1400 985 0.011 0.005 0.77 0.03 2.836 3.755 0.477 0.034 0.24 0.01 3.7 0.6 0.18 0.02 0.48 0.07 3.1+2.9 2.9+2.7 167.4 604.4 0.029 0.007 1.08 0.24 4.496 4.977 0.883 0.163 0.4 0.04 4.8 1.1 0.29 0.07 0.33 0.04 Crushed % % % Crushed % % % Size (mm) Size WeightSinks Size (mm) Size weight Size (mm) Size weight Size (mm) Size weight Uncrushed % % % Uncrushed % % % Head GradeHead 17618.2 0.036 0.757 4.235 1.167 0.031 3.187 0.20 Head GradeHead 10635 0.038 0.801 3.853 1.193 0.034 3.357 0.253 0 Head GradeHead 7780 0.044 0.971 2.765 1.498 0.091 4.220 0.202 0. Head GradeHead 1582.1 0.047 0.815 3.531 1.429 0.082 3.663 0.388
DMS Test Results
180 t t /mt /mt Pt Pt Pt Pt 338 021 481 .426 Assay Assay Assay Assay % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g % g/mt g/mt g/mt g Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd Tails 8000 0.011 0.16 4.076 0.182 0.03 0.6 0.07 0.21 Tails 6000 0.02 0.46 3.536 0.486 0.02 1.8 0.13 0.23 Tails 1200 0.025 0.15 4.616 0.733 0.05 0.6 0.36 0.16 Tails 800 0.022 0.31 3.974 0.63 0.05 1.2 0.13 0.13 Conc 11200 0.054 1.04 3.934 1.839 0.04 4.5 0.46 0.58 Conc 2800 0.08 0.75 2.777 2.445 0.04 4.2 0.41 0.57 Conc 8000 0.073 0.47 3.125 2.239 0.64 4.3 0.51 1.15 Conc 600 0.105 0.99 1.923 3.329 1.65 11.3 0.81 0.95 Tails2 Tails2 Tails2 Tails2 75+25 (gm) Co Cu Mg 75+25 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg Crushed % % % Crushed % % % Size (mm) Size Weight Size (mm) Size weight Size (mm) Size weight Size (mm) Size weight Uncrushed % % % Uncrushed % % % Head GradeHead 19200 0.036 0.673 3.993 1.149 0.036 2.875 0.298 0 Head GradeHead 8800 0.039 0.552 3.295 1.109 0.026 2.564 0.219 0. Head GradeHead 9200 0.067 0.428 3.319 2.043 0.563 3.817 0.490 1. Head GradeHead 1400 0.058 0.601 3.095 1.787 0.736 5.529 0.421 0. Sorting TestResults
181 0.526 0.271 5.049 0.140 g/mt g/mt g/mt g/mt iAu Ag Pd Pt Assay 1.676 3.189 1.302 0.053 % % % % (gm) Co Cu Mg N 7800 weight Crush 6200 0.051 1.22 3.22 1.643 0.05 4.7 0.25 0.54 Uncrush 1600 0.059 1.62 3.068 1.806 0.49 6.4 0.35 0.47 Fines Fines 6.7 Head Grade Head
182 Montcalm East t t /mt /mt Pt Pt Pt Pt 0.000 0.000 0.008 0 0.000 Assay Assay Assay Assay % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g % g/mt g/mt g/mt g Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd 2.8 802.3 0.008 0.16 6.297 0.151 0.01 0.7 0 0 2.8 525 0.004 0.09 5.388 0.071 0.02 0.9 0 0 2.8 204.7 0.005 0.18 5.192 0.104 0.04 0.6 0.01 0 2.8 135.9 0.007 0.17 4.937 0.169 0.01 0.5 0 0 +3.1 3526.3 0.121 0.65 2.095 3.773 0.03 0.7 0 0 +3.1 13241.6 0.084 0.78 3.781 2.685 0.07 2.2 0 0 +3.1 4841.8 0.044 1.54 4.373 1.658 0.07 3.3 0 0 +3.1 597 0.099 0.74 3.212 2.852 0.03 2.3 0 0 75+25 (gm) Co Cu Mg 75+25 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg Crushed % % % Crushed % % % 3.1+2.95 2.95+2.8 1776.7 2539.1 0.036 0.011 1.27 0.25 5.372 6.211 1.132 0.221 0.04 0 3.1 0.7 0 0.02 0 0 3.1+2.95 2.95+2.8 2538 5914.2 0.025 0.007 0.56 0.1 6.141 6.424 0.612 0.107 0.02 0.02 1.3 0.4 0 0 0 0 3.1+2.95 2.95+2.8 3585 963.2 0.03 0.011 0.31 0.33 7.274 5.638 0.699 0.274 0.01 0.02 1.1 0.8 0 0 0 0 3.1+2.95 2.95+2.8 197.3 313.9 0.038 0.012 0.94 0.8 5.376 6.103 1.607 0.318 0.04 0.04 2.7 2.3 0 0 0 0.03 Size (mm) Size weight Size (mm) Size WeightSinks Size (mm) Size weight Size (mm) Size weight Uncrushed % % % Uncrushed % % % Head GradeHead 8644.4 0.061 0.614 4.368 1.851 0.021 1.193 0.006 Head GradeHead 22218.8 0.055 0.558 4.792 1.700 0.050 1.587 0.00 Head GradeHead 9594.7 0.035 0.930 5.601 1.128 0.042 2.169 0.000 Head GradeHead 1244.1 0.057 0.725 4.473 1.722 0.032 2.167 0.000
DMS Test Results
183 t t /mt /mt Pt Pt Pt Pt 056 0.000 0 0.000 Assay Assay Assay Assay % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g % g/mt g/mt g/mt g Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd Tails 653.9 0.009 0.08 6.579 0.157 0.01 0.4 0.01 0 Tails 3323.1 0.037 0.68 5.311 0.972 0.08 2.1 0 0.1 Tails 414.2 0.003 0.01 6.027 0.037 0.01 0.4 0 0 Tails Conc 18715.3 0.069 0.49 4.073 2.092 0.06 1.6 0 0 Conc 2624.9 0.086 0.73 3.3 2.58 0.04 2 0 0 Conc 7187.2 0.051 0.96 4.602 1.875 0.04 2.7 0.01 0 Conc Tails2 4169 0.011 0.11 6.261 0.162 0.01 0.6 0 0 Tails2 Tails2 1746 0.009 0.35 6.127 0.241 0.09 1 0.02 0 Tails2 75+25 (gm) Co Cu Mg 75+25 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg Crushed % % % Crushed % % % Uncrushed % % % Uncrushed % % % Head GradeHead 23538.2 0.057 0.411 4.530 1.696 0.050 1.390 0.00 Head GradeHead 5948 0.059 0.702 4.424 1.682 0.062 2.056 0.000 0. Head GradeHead 9347.4 0.041 0.804 4.950 1.488 0.048 2.281 0.011 Head GradeHead 0 Sorting Test Results Size (mm) Size Weight Size (mm) Size weight Size (mm) Size weight Size (mm) Size weight
184 0.000 0.018 1.890 0.380 g/mt g/mt g/mt g/mt iAu Ag Pd Pt Assay 1.847 4.499 0.648 0.061 % % % % (gm) Co Cu Mg N 6010 weight Crush 5410 0.06 0.64 4.501 1.825 0.4 1.9 0.02 0 Uncrush 600 0.067 0.72 4.482 2.043 0.2 1.8 0 0 Fines Fines 6.7 Head Grade Head
185 Montcalm West t t /mt /mt Pt Pt Pt Pt 000 .000 0 0.000 0 0.000 Assay Assay Assay Assay % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g % g/mt g/mt g/mt g Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd 2.8 206.1 0.005 0.05 5.526 0.101 0 0.7 0 0 2.8 4573.6 0.004 0.03 4.838 0.044 0 0.2 0 0 2.8 2.8 4733.1 0.004 0.04 2.895 0.036 0 0.5 0 0 +3.1 110.6 0.159 0.74 2.12 3.473 0.04 2.7 0 0 +3.1 415.6 0.05 0.46 6.069 1.406 0.03 2.1 0 0 +3.1 1330.6 0.119 0.85 3.506 3.274 0.04 3 0 0 +3.1 1816.1 0.044 0.56 6.014 1.403 0.04 2.3 0 0 75+25 (gm) Co Cu Mg 75+25 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg Crushed% % % Crushed% % % 3.1+2.95 2.95+2.8 61.4 285.3 0.045 0.016 0.46 0.2 4.881 7.013 1.284 0.407 0.09 0.02 18.7 1 0 0 0 0 3.1+2.95 2.95+2.8 889.9 4858.7 0.035 0.009 0.39 0.07 6.047 6.721 1.127 0.171 0.05 0 2 0.4 0 0 0 0 3.1+2.95 2.95+2.8 651.3 1904.1 0.018 0.01 1.35 0.14 5.742 8.26 0.659 0.228 0.39 0.04 7.2 0.8 0 0 0 0 3.1+2.95 2.95+2.8 1505.1 10587.4 0.029 0.007 0.23 0.11 5.689 5.857 0.907 0.145 0.03 0.01 1 0.7 0 0 0 0 Size (mm)Size weight Size (mm)Size weight Size (mm)Size Weight Sinks Size (mm)Size weight Uncrushed% % % Uncrushed% % % Head Grade 663.4 0.039 0.267 5.538 0.904 0.024 2.828 0.000 0 Head Grade 10737.8 0.011 0.095 5.838 0.244 0.005 0.513 0.00 Head Grade 3886 0.049 0.586 6.210 1.343 0.099 2.626 0.000 0. Head Grade 18641.7 0.012 0.146 5.107 0.301 0.012 0.829 0.00
DMS Test Results
186 t t /mt /mt Pt Pt Pt Pt .000 .000 0.000 0 0.000 0.000 0.994 0.015 Assay Assay Assay Assay % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g/m % g/mt g/mt g/mt g % g/mt g/mt g/mt g Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd Ni Au Ag Pd 0.420 7.001 0.214 0.013 4301.6 Tails 787.4 0.032 0.44 6.2 0.733 0.05 1.8 0 0 Tails 10883.8 0.013 0.14 5.817 0.295 0.02 0.5 0 0 Tails 2165.9 0.007 0.05 7.207 0.12 0 0.2 0 0 Tails 15454.7 0.005 0.04 5.801 0.066 0.01 0.2 0 0 Conc 108 0.101 1.11 3.374 3.135 0.1 2.9 0 0 Conc 62.2 0.05 0.57 5.858 1.597 0.06 2.4 0 0 Conc 2135.7 0.02 0.38 6.793 0.725 0.03 1.8 0 0 Conc 6255.7 0.022 0.31 6.055 0.658 0.05 1.5 0 0 Tails 2 Tails 2 Tails 2 Tails 2 1050 0.016 0.21 6.46 0.542 0.01 0.7 0 0 75+25 (gm) Co Cu Mg 75+25 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg 25+6.7 (gm) Co Cu Mg Crushed% % % Crushed% % % Size (mm)Size weight Size (mm)Size Weight (mm)Size weight Size (mm)Size weight Uncrushed% % % Uncrushed% % % Head Grade 895.4 0.040 0.521 5.859 1.023 0.056 1.933 0.000 0 Head Grade 10946 0.013 0.142 5.817 0.302 0.020 0.511 0.000 0 Head Grade Head Grade 22760.4 0.010 0.122 5.901 0.251 0.021 0.580 0.00 Sorting Test Results
187 0.000 0.000 0.910 g/mt g/mt g/mt g/mt 0.081 i Au Ag Pd Pt Assay 0.501 5.731 0.217 0.020 % % % % (gm) Co Cu Mg N 8330 weight Crush 6870 0.016 0.17 5.803 0.379 0.09 0.7 0 0 Uncrush 1460 0.04 0.44 5.391 1.076 0.04 1.9 0 0 Fines 6.7 Head Grade
188 Appendix 4: Physical and Geotechnical Properties of Fill Mixes 18 5.09 6.68 5.09 26.71 % 4.75mm % Cement / % Cement 0.32 0.66 0.14 0.83 UCS UCS/ %Cement 0.12 0.09 0.14 0.10 Porosity Binder Binder / 0.58 0.58 0.07 0.58 (von Mises) (von 1.60 1.40 0.79 1.75 MPa 0.74 0.30 0.67 0.28 0.43 0.23 0.40 0.22 Poured Ratio % 3.4 9.4 8.1 % 18.7 41.4 82.3 41.4 2.3 mm 20.0 18.2 18.2 80 80 % Passing 8.8 6.0 56.9 0.41 0.71 1.24 0.58 0.12 0.25 8.79 mm 37.5 37.5 37.5 37.5 37.5 15.0 38.4 0.41 0.69 1.47 0.58 0.12 0.29 13.03 Top Size4.75 mm .02 mm Porosity Void UCS tau' u 8 37.5 22.0 18.4 0.38 0.62 2.11 0.58 0.13 0.42 27.25 C 6.0 6.0 6.0 6.0 u C 29.3 6.8 26.5 7.0 64.9 7.1 0.20 0.25 0.75 0.58 0.09 0.41 2.80 95.3 4.7 37.5 11.3 54.6 7.6 0.22 0.28 0.54 0.58 0.09 0.27 3.68 6.0 6.5 70.5 3.8 18.8 5.9 68.0 3.1 0.23 0.29 1.66 0.58 0.09 0.85 2.87 217.2 455.3 440.1 6.8 37.5 20.7 38.8 6.5 0.19 0.23 0.63 0.58 0.09 0.39 4.18 Particle Size Particle 0 98.5 4.7 37.5 11.3 54.6 3.2 0.22 0.28 1.36 0.58 0.09 0.67 3.68 1 71.1 6.8 26.5 7.0 64.9 3.0 0.20 0.25 1.55 0.58 0.09 0.85 2.80 9 6.6 3.8 18.8 1.4 90.1 9.2 0.41 0.69 0.75 0.17 0.13 0.14 6.10 c 08 24.7 6.8 26.5 4.4 81.1 7.1 0.31 0.45 0.98 0.26 0.13 0.24 4.91 45 238.2 6.8 37.5 20.7 38.8 2.7 0.19 0.23 0.57 0.58 0.09 0.35 4. 028 6.6 6.8 37.5 2.6 81.9 9.3 0.40 0.66 0.92 0.09 0.14 0.17 6.68 C 3.920 0.315 0.029 0.500 Curvature Uniformity Coefficent of Coefficent of Coefficient Cu Coarse % 2.0 7.3 6.8 7.4 Height Change in in Change 1.0 4.9 4.8 4.9 Cement Water / % 4.8 9.4 9.4 21.0 Water Content % 5.0 2.1 5.5 2.1 1.7 2.3 2.0 2.5 % 0.0 28.9 75.1 28.9 Tailings Full Full Full Type Cycloned Cycloned Cycloned % 95.0 69.0 19.4 76.3 CT 1:3 CT 1:7 CT Full 1:3 Full 1:7 Full CTMax RockFill Full Max Full Ore Body Ore Mix Rejects SG Cement TL TL Zone 1 TL Zone 1 Max Full TL Zone 1 1:3 Full 1:7 Full Full Full Full TL TL Zone 1 RockFill 95.0 0.0 1.7 5.0 4.8 1.0 1.1 1.689 3.8 3.8 1 TL TL Zone 1 TL Zone 1 CTMax TL Zone 1 71.51:3 CT Cycloned1:7 CT 26.6 2.4 19.4 1.9 Cycloned Cycloned 75.1 8.7 2.1 4.9 5.5 21.0 0.0 4.8 0.160 4.4 0.02 TL TL Zone 2 TL Zone 2 TL Zone 2 TL Zone 2 TL Zone 2 TL Zone 2 TL Zone 2 Craig 8112Craig 8112Craig RockFill CTMax 71.5 Cycloned 26.7 2.1 1.8 8.2 4.9 5.3 0.14 Craig 8112Craig 1:3 CT 35.9 8112Craig Cycloned 60.2 2.0 8112Craig 1:3 Full 4.0 1:7 Full 16.3 4.9 Full 7.3 Full 0.0 Craig 8112Craig 8112Craig 1:7 CT Max Full 71.5 Cycloned Full 26.7 2.3 1.8 8.2 4.9 2.0 0.140 1 TLFootwall TLFootwall 1:3 CT 1:7 CT TLFootwall 19.8 Cycloned Cycloned TLFootwall1:3 Full 74.7 2.0 1:7 Full 5.5 20.9 Full 4.8 Full 9.8 0. TLFootwall Max Full 76.3 Full 22.1 2.2 1.6 7.3 4.9 0.0 0.344 TLFootwall RockFillTLFootwall 95.0 CTMax 76.3 Cycloned 22.1 0.0 2.1 1.8 1.6 5.0 7.3 4.8 4.9 1.0 0.0 1.2 0.3 3.379 6.8 6. Craig LGBX Craig RockFillLGBX Craig 95.0 CTMax LGBX Craig 70.7LGBX Craig 1:3 CT CyclonedLGBX Craig 1:7 CT 27.3 0.0 2.2 Max Full LGBX Craig 1.8 2.0 Cycloned 70.7 1:3 Full LGBX Craig 5.0 8.9 Cycloned 1:7 Full Full 4.8 4.9 27.3 1.0 2.2 0.0 Full 2.0 2.8 Full 8.9 0.12 2.227 4.9 1.0 4.7 0.120 4.7 1
189 1 6 7 24 86 81 .92 .33 .86 .65 6.76 % 4.75mm % Cement / % Cement 8 6.33 7 6.69 0 6.81 9 6.92 0 4.70 9 5.95 .50 4.93 0.13 UCS UCS/ %Cement 2 0.19 6.61 0.14 Porosity Binder Binder / 0.05 0.70 0.66 0.40 Poured Ratio 22.17 80.63 4.56 80 80 % Passing 7.5 20.0 22.0 0.42 0.71 1.78 0.58 0.12 0.36 22.77 mm mm % % MPa Mises) (von 37.5 21.0 17.7 0.49 0.95 2.11 0.58 0.10 0.42 28.19 37.50 Top Size4.75 mm .02 mm Porosity Void UCS tau' u C 5.1 u 9 4.9 37.5 21.0 11.6 0.51 1.02 1.43 0.58 0.10 0.29 43.13 5.1 6.0 37.5 17.9 41.5 7.2 0.23 0.29 1.17 0.58 0.09 0.59 4.80 .6 3.6 37.5 22.0 8.9 0.49 0.94 3.41 0.58 0.10 0.68 56.07 .2 5.2 35.16 18.37 24.06 0.00 0.44 0.80 1.89 0.58 0.11 0.38 28. C 40.8 5.7 37.5 18.7 43.5 9.0 0.23 0.29 1.33 0.58 0.11 0.54 5.61 23.7 24.0 6.0 37.5 3.7 81.6 22.0 0.40 0.67 0.83 0.05 0.14 0.15 6.67 23.6 5.7 37.5 3.9 81.0 22.2 0.39 0.64 0.70 0.05 0.14 6.76 247.1 5.7 37.5 13.6 66.0 16.9 0.35 0.55 0.74 0.12 0.12 0.17 6.5 186.8 6.0248.3 37.5 17.9 6.0 41.5 3.0 37.5 0.23 13.0 0.29 67.3 1.44 16.7 0.35 0.58 0.55 0.09 0.96 0.72 0.11 4.80 0.12 0.22 6.3 Particle Size Particle 8 200.2 5.7 37.5 18.7 43.5 3.8 0.25 0.34 1.32 0.58 0.10 0.54 5.6 38 78.4 6.6 5.7 5.7 37.5 13.6 37.5 66.0 3.9 7.0 81.0 0.35 9.2 0.55 0.40 0.81 0.67 0.70 0.11 0.07 0.12 0.19 0.14 0.13 6.56 6.76 1 191.7 5.4 34.20 15.83 47.49 3.14 0.23 0.30 1.24 0.58 0.09 0.6 39 78.8 5.1 6.0 6.0 37.5 13.0 37.5 67.3 3.7 7.0 81.6 0.35 9.2 0.55 0.40 0.62 0.67 0.76 0.09 0.07 0.12 0.14 0.14 0.14 6.38 6.67 c 04 60.7 6.1 33.83 10.34 71.48 7.04 0.34 0.51 0.80 0.15 0.12 0.1 30 6.4 5.3 34.83 3.55 82.30 9.26 0.40 0.67 0.77 0.09 0.14 0.14 6 003 720.6 4.9 37.5 18.6 36.7 9.7 0.22 0.29 1.22 0.58 0.11 0.48 6 006 709.7 3.6 37.5 24.1 38.0 9.2 0.24 0.32 1.16 0.58 0.10 0.48 6 196 429.5 5.5 36.13 17.06 44.93 8.05 0.22 0.28 1.07 0.58 0.10 0 C .003 228.8 4.9 37.5 15.6 64.0 17.1 0.33 0.50 0.74 0.07 0.13 0.1 .001 232.7 3.6 37.5 16.2 63.1 17.2 0.35 0.55 0.84 0.20 0.12 0.2 .007 23.6 3.6 37.5 6.1 79.4 22.3 0.40 0.67 0.55 0.06 0.14 0.10 6 .002 239.2 5.0 37.50 14.58 65.10 16.98 0.35 0.53 0.82 0.13 0.1 0.004 326.5 3.6 37.5 24.1 38.0 2.2 0.25 0.34 0.91 0.58 0.10 0.3 0.028 6.6 3.6 37.5 6.1 79.4 9.3 0.40 0.67 0.68 0.07 0.14 0.13 6. 0.007 0.029 6.6 4.9 37.5 4.9 79.9 9.3 0.40 0.67 0.83 0.11 0.14 0.15 6. 0.003 315.9 4.9 37.5 18.6 36.7 4.0 0.26 0.36 1.00 0.58 0.10 0.3 Curvature Uniformity 5.40 Height Change in in Change 4.37 Cement Water / 19.26 Water Content 5.45 2.16 74.46 Tailings Full Type Pouredof Coefficent of Coefficient Cu Coarse %%%%% 20.08 Full 1:7 Full Average Average CTMax 69.13 Sand 28.77 2.26 2.10 9.27 4.88 3.89 0.17 Average1:3 CT 36.52 Sand 59.28 2.05 4.20 17.05 4.89 6.83 0.0 Average RockFill 95.00 11.88 1.73 5.00 4.76 1.00 5.62 4.885 5 AverageAverage1:7 CT Max Full 19.87 69.74 Sand Full 74.66 1.98 29.05 5.47 2.34 2.12 20.92 9.34 4.84 4.87 5.78 2.48 0.0 0. Average 1:3 Full 37.04 Full 58.66 2.20 4.30 15.81 4.37 3.31 0 Fraser Fraser Ni 1:7 Full 20.25 Full 74.3 2.1 5.4 19.2 4.4 4.9 0.007 Fraser Fraser Ni Fraser Ni RockFill Fraser Ni CTMax 95.0 Fraser Ni1:3 CT 71.0 Fraser Ni Cycloned1:7 CT 37.1 27.0 Fraser Ni Max Full 2.4 0.0 Cycloned 20.3 58.6 1.7 71.0 2.0 1:3 Full Cycloned 2.1 74.3 5.0 37.20 8.8 4.3 1.9 Full 4.8 5.4 27.0 17.3 Full 4.8 2.4 1.0 58.5 20.8 4.9 2.0 4.9 2.2 4.8 0.9 4.3 4.9 8.8 0.441 4.1 15.8 4.8 2.880 0.00 4.4 1.0 0.03 3.4 6.0 0.440 0.001 6.0 34 3 Fraser Cu Fraser Cu Fraser RockFill CTMax 95.0 64.3 Cycloned 33.3 2.2 0.0 1.7 2.4 5.0 10.7 4.8 4.9 1.0 5.4 12.8 0.00 3.018 5.7 5.7 Ore Body Ore Mix Rejects SG Cement Fraser Cu Fraser Cu Fraser 1:3 CT Cu Fraser 1:7 CT 36.6Cu Fraser Max Full Cycloned 19.8 59.1 64.3 1:3 Full Cycloned 2.1 74.7 36.57 4.3 1.9 Full 5.5 33.3 17.5 Full 2.4 59.1 20.9 4.9 2.4 2.2 4.8 4.3 10.7 8.3 2.9 15.9 4.9 0.00 4.4 3.0 0.02 4.4 0.014 0.001 4 Fraser Cu Fraser 1:7 Full 19.80 Full 74.7 2.1 5.5 19.3 4.4 6.0 0.007 Montcalm EastMontcalm 1:7 Full Full Montcalm EastMontcalm 1:3 Full 37.28 Full 58.4 2.2 4.3 15.8 4.4 3.0 0 Montcalm EastMontcalm Max Full 63.0 Full 34.5 2.4 2.5 11.0 4.9 2.6 0. Montcalm EastMontcalm EastMontcalm 1:3 CT 1:7 CT 20.3 Cycloned Cycloned 74.3 2.0 5.4 20.8 4.8 7.3 Montcalm EastMontcalm RockFill EastMontcalm CTMax 95.0 63.0 Cycloned 34.5 2.3 95.0 1.6 2.5 5.0 11.0 4.8 4.9 1.0 7.3 8.3 8.963 4. Montcalm WestMontcalm RockFill WestMontcalm CTMax 95.0 64.8 Cycloned 32.8 2.3 0.0 1.8 2.4 5.0 10.5 4.8 4.9 1.0 4.9 15.8 13.000 3 Montcalm WestMontcalm 1:3 CT WestMontcalm 1:7 CT 20.2 Cycloned Cycloned 74.4 2.0 5.4 20.9 4.8 5.1 Montcalm WestMontcalm Max Full 64.8 Full 32.8 2.4 2.4 10.5 4.9 3.0 0. Montcalm WestMontcalm 1:3 Full 37.12 Full 58.6 2.2 4.3 15.8 4.4 2.4 0 Montcalm WestMontcalm 1:7 Full 20.20 Full 74.4 2.2 5.4 19.2 4.4 5.4 0 190