ADVANCED BENEFICIATION of BASTNAESITE ORE THROUGH CENTRIFUGAL CONCENTRATION and FROTH FLOTATION by Doug Schriner

ADVANCED BENEFICIATION OF BASTNAESITE ORE THROUGH

CENTRIFUGAL CONCENTRATION AND FROTH FLOTATION

by Doug Schriner

A thesis submitted to the Faculty and the Board of Trustees of the Colorado School of Mines in partial fulfillment of the requirements for the degree of Master of Science (Metallurgical and Materials Engineering).

Golden, Colorado Date ______

Signed: ______Doug Schriner

Signed: ______Dr. Corby Anderson Thesis Advisor

Golden, Colorado Date ______

Signed: ______Dr. Ivar Reimanis Professor and Head Department of Metallurgical and Materials Engineering

ABSTRACT

Gravity separation and flotation studies have been conducted on Molycorp bastnaesite ore in order to determine if new beneficiation schemes present a more selective and more economical alternative than that which is currently employed at Mountain Pass. Literature on bastnaesite, monazite, barite, and calcite flotation and gravity concentration principles was surveyed. Flotation reagent additions were determined using components that have shown preferential floatability of bastnaesite and monazite over the gangue minerals. Hallimond Tube microflotation tests were performed on crushed and ground ore samples. Heavy liquid separation with sodium polytungstate was used to investigate the effectiveness of gravity separation on the ore. Shaking table and Falcon concentrator tests were performed to gravity concentrate the ore. A gravity-concentrated feed was floated and compared with a non-concentrated ore feed to illustrate the benefit of preconcentration. An economic analysis was generated for flotation plants operating with and without gravity preconcentration that would sell products with two distinct grades and recoveries.

Qualitative microflotation tests produced little selective separation of the rare earth minerals (bastnaesite, parisite, and monazite) from the gangue (calcite, barite, dolomite, and quartz). Heavy liquid tests illustrated the sink/float behavior of the minerals at different specific gravities of separation. Their results suggest that at higher specific gravities the calcite floats while the bastnaesite and barite sink. Shaking table tests showed potential to effect such a separation, but optimum conditions were not determined. A Falcon centrifugal concentrator was used to carry out tests according to a Design of Experiments matrix generated with Stat Ease Design Expert 9. The best conditions from those trials were determined, and the tests were repeated to verify the desirability of those parameters. Bench flotation was then used to compare the standard feed at plant conditions to a feed consisting of the blended gravity concentrates. The flotation results showed that the preconcentrated feed outperformed the typical plant feed. Economic analysis of a plant with and without gravity preconcentration shows that gravity preconcentration, although more capital-intensive, will yield a higher annual profit and a better 10-year net present value.

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TABLE OF CONTENTS

ABSTRACT ...... iii

LIST OF FIGURES ...... vi

LIST OF TABLES ...... xi

CHAPTER 1: INTRODUCTION ...... 1

CHAPTER 2: LITERATURE REVIEW ...... 9

2.1 Flotation Surface Chemistry and Analysis Techniques ...... 9

2.1.1 Hydrophobicity and Contact Angle ...... 10

2.1.2 Hydrolysis Reactions ...... 11

2.1.3 Zeta Potential and Point of Zero Charge ...... 13

2.1.4 Adsorption Density ...... 17

2.1.5 Hallimond Tube Flotation...... 19

2.1.6 Temperature Effects ...... 19

2.2 Minerals ...... 21

2.2.1 Bastnaesite ...... 21

2.2.2 Monazite ...... 22

2.2.3 Calcite and Barite ...... 22

2.3 Reagents ...... 23

2.3.1 Collectors ...... 23

2.3.1.1 Fatty Acids ...... 23

2.3.1.2 Hydroxamates ...... 26

2.3.2 Modifiers and Depressants ...... 29

2.3.2.1 Soda Ash ...... 29

2.3.2.2 Lignin Sulfonate ...... 29

2.3.2.3 Metal Salts ...... 29

2.3.2.4 Sodium Silicate ...... 31

2.4 Gravity Concentration ...... 33

2.4.1 Heavy Liquid Separation ...... 37

2.4.2 Shaking Tables ...... 37

2.4.3 Knelson and Falcon Concentrators ...... 39

CHAPTER 3: EXPERIMENTAL METHODS ...... 43

3.1 Characterization and Mineralogy Procedures ...... 43

3.2 Microflotation ...... 43

3.3 Magnetic Separation ...... 46

3.4 Gravity Concentration ...... 46

3.5 Bench Flotation ...... 48

CHAPTER 4: RESULTS AND DISCUSSION ...... 49

4.1 Characterization and Mineralogy ...... 49

4.2 Microflotation ...... 55

4.3 Magnetic Separation ...... 61

4.4 Gravity Concentration ...... 62

4.5 Bench Flotation ...... 70

CHAPTER 5: ECONOMIC ANALYSIS ...... 71

CHAPTER 6: CONCLUSIONS ...... 76

CHAPTER 7: SUGGESTIONS FOR FUTURE WORK ...... 78

REFERENCES ...... 79

APPENDIX A: FIRST LOT MINERALOGY ...... 83

APPENDIX B: SECOND LOT MINERALOGY ...... 91

APPENDIX C: EXPERIMENTAL DATA ...... 98

APPENDIX D: ECONOMIC ESTIMATES ...... 101

LIST OF FIGURES

Figure 1.1. Beneficiation Flowsheet at Mountain Pass (as of 1991) [2] ...... 2

Figure 1.2. Molycorp REO Flotation Circuit ...... 3

Figure 1.3. REO Export Prices from China. Data retrieved from www.metal- pages.com ...... 4

Figure 1.4. Sichuan Mianning REE Deposit Flowsheet. [10] ...... 7

Figure 1.5. Egyptian Beach Sand Beneficiation Flowsheet. The first percentage given relates to the original weight and the second represents the assay of monazite. [12] ...... 8

Figure 2.1. Visual Representation of Young's Equation, showing the contact angle of a hydrophobic (left) and hydrophilic (right) surface. [19] ...... 10

Figure 2.2. Illustration of the geometry used to determine the contact angle of water on carbon. [19] ...... 11

Figure 2.3. Aqueous Equilibria of cerium species. [22] ...... 12

Figure 2.4. Schematic of electrical double layer (from Somasundaran 1975). [23] ...... 13

Figure 2.5. Zeta Potential of bastnaesite and monazite. (Luo Jiake 1984, from [16]) ...... 15

Figure 2.6. Zeta potential of barite, bastnaesite, and calcite in pure water. (Smith 1986, from [16])...... 16

Figure 2.7. Oleate adsorption onto bastnaesite as a function of pH with and without pre-boiling. (Smith 1986 from [16] ...... 17

Figure 2.8. Adsorption isotherm of oleate on calcite at pH 9.3. [21] ...... 18

Figure 2.9. Adsorption Density of hydroxamate on barite, calcite, and bastnaesite at 21°C and pH 9.3. [22] ...... 18

Figure 2.10. Schematic Drawing of a Modified Hallimond Tube. [31] ...... 20

Figure 2.11. Effect of pulp temperature on oleate flotation...... 20

Figure 2.12. The flotation recoveries of Mountain Pass ore as a function of conditioning temperature for fatty acids (left) and hydroxamate (right) collectors. [22] ...... 21

Figure 2.13. Calcite recovery as a function of fatty acid addition for various fatty acids at pH 9.7. [34] ...... 24

Figure 2.14. Calcite recovery as a function of pH for various fatty acids at a constant collector addition of 10-3 mol/liter. [34] ...... 24

Figure 2.15. Structure of oleic acid ...... 24

Figure 2.16. Bastnaesite recovery by sodium oleate as a function of pH. [26] ...... 25

Figure 2.17. Oleate concentration required to float salt-type minerals.[21] ...... 26

Figure 2.18. A mechanism of hydroxamate adsorption to an ion on a mineral surface. [20] ...... 26

Figure 2.19. The structure of potassium octyl hydroxamate. [22] ...... 27

Figure 2.20. Modified hydroxamic acid chelating surface cerium(III). [37] ...... 27

Figure 2.21. Hallimond tube flotation results on bastnaesite with K-octyl hydroxamate as collector. [22] ...... 28

Figure 2.22. Bastnaesite ore recovery (as total weight, including gangue, recovered) by K-octyl hydroxamate as a function of pH. [22] ...... 28

Figure 2.23: Bastnaesite recovery as a function of ammonium lignin sulfonate concentration. [26] ...... 30

Figure 2.24. Recovery of calcite as a function of increased concentration of metal salts. [33] ...... 30

Figure 2.25: Effect of sodium metasilicate for oleate (top) and hydroxamate (bottom) flotation. [38] ...... 31

Figure 2.26. Adsorption and Floatability of sodium oleate and sodium silicate on calcite and fluorite as a function of sodium silicate concentration. [21] ...... 32

Figure 2.27. Floatability of different minerals versus Na2SiO3. (Luo Jiake 1984, from [16] ...... 32

Figure 2.28. Applicability of Beneficiation Equipment to a given feed size. [41] ..... 35

Figure 2.29. Effect of contaminating slimes and suspension viscosity on the cleaning of Roslyn coal. [42] ...... 36

Figure 2.30. XRD Patterns of Molycorp Bastnaesite size fractions upgraded with a Frantz Isodynamic Separator and reference minerals. All

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diffraction peaks have been normalized to the maximum peak intensity for each pattern. [40] ...... 36

Figure 2.31. Density of aqueous sodium polytungstate solution as a function of mass percent. [43] ...... 37

Figure 2.32. Deister shaking table deck. Dark circles represent low-s.g. particles. White circles represent high-s.g. particles. The diameter of the circles corresponds of the particle size. [45] ...... 38

Figure 2.33. Fluid Motion over riffles on a Deister deck. [45] ...... 38

Figure 2.34. Wilfley shaking table ...... 39

Figure 2.35. Knelson Concentrator schematic. [46] ...... 40

Figure 2.36. Cumulative gold retained as a function of particle size. [47] ...... 41

Figure 2.37. Cross-Section of a Falcon Concentrator ...... 42

Figure 3.1. Modified Hallimond Tube used for microflotation tests...... 45

Figure 4.1. Particle size distributions for ground ore samples ...... 49

Figure 4.2. False-Color image of bastnaesite ore +50 mesh size fraction. Values represent surface area percentages...... 52

Figure 4.3. False-Color image of 90-minute ground bastnaesite ore 200 x 400 mesh size fraction. Values represent surface area percentages...... 52

Figure 4.4. Cumulative REE mineral recovery by liberation class from MLA...... 54

Figure 4.5. Molycorp ore zeta potential ...... 54

Figure 4.6. Elemental recovery as a function of collector concentration ...... 56

Figure 4.7. Concentrate grade as a function of collector concentration...... 56

Figure 4.8. Elemental recovery as a function of ammonium lignin sulfonate concentration...... 57

Figure 4.9. Concentrate grade as a function of ammonium lignin sulfonate concentration...... 57

Figure 4.10. Elemental recovery as a function of sodium silicate concentration. .... 58

Figure 4.11. Concentrate grade as a function of sodium silicate concentration...... 59

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Figure 4.12. Elemental recovery as a function of copper (II) nitrate concentration...... 59

Figure 4.13. Concentrate grade as a function of copper (II) nitrate concentration. . 60

Figure 4.14. WHIMS concentrate TREE grade (left)and recovery (right)...... 62

Figure 4.15. Barium distribution between WHIMS magnetic and non-magnetic fractions...... 62

Figure 4.16. Calcium recoveries to HLA products...... 63

Figure 4.17. Mass recovery as a function of fluid density...... 64

Figure 4.18. Shaking Table Streams ...... 65

Figure 4.19. Gravity concentration circuit...... 67

Figure 4.20. Desirability Surface for Falcon tests with 50-micron feed size...... 69

Figure 4.21. TREE Grade vs Recovery of Falcon Concentrates...... 70

Figure 5.1. Sensitivity Analysis for the gravity-flotation concentrator...... 73

Figure A.1. Cumulative particle size analysis of ground Molycorp bastnaesite ore...... 83

Figure A.2. Liberation of Bastnaesite according to size fraction. Left: +100 mesh, Center: 100 x 325 mesh, Right: 325 x 400 mesh...... 85

Figure A.3. False-color image showing particle association of the +100 mesh size fraction...... 85

Figure A.4. False-color image showing particle association of the 100 x 325 mesh size fraction...... 86

Figure A.5. False-color image showing particle association of the 325 x 400 mesh size fraction...... 86

Figure A.6. Measured and WPPF-calculated diffractograms and residual plots for the bastnaesite ore...... 87

Figure A.7. Bastnaesite ore diffractogram with candidate phases...... 87

Figure A.8. QEMSCAN mineral mass abundance in each size fraction...... 90

Figure B.1. Mineral locking for bastnaesite...... 93

Figure B.2. Mineral locking for parisite...... 93

Figure B.3. Mineral locking for monazite...... 94

Figure B.4. Mineral locking for calcite...... 94

Figure B.5. False-color image showing particle association of the +100 mesh size fraction...... 95

Figure B.6. False-color image showing particle association of the +100 mesh size fraction...... 95

Figure B.7. False-color image showing particle association of the +100 mesh size fraction...... 96

Figure B.8. Measured and WPPF-calculated diffractograms and residual plot for the original bastnaesite ore...... 96

Figure B.9. Bastnaesite ore diffractogram with candidate phases...... 97

Figure D.1. Sensitivity Analysis for the flotation concentrator...... 101

LIST OF TABLES

Table 1.1. Flotation Conditioning Steps at Mountain Pass. [3] ...... 2

Table 1.2. Characteristics of RER scrubbing tests. [14] ...... 7

Table 2.1. Bastnaesite Points of Zero Charge (PZC). [24] ...... 14

Table 2.2. Monazite Points of Zero Charge (PZC). [24] ...... 14

Table 2.3. Points of Zero Charge (PZC) of alkaline-earth semisoluble salt minerals. [24] ...... 15

Table 2.4. Specific Gravity (SG) and Concentration Criterion (CC) for major components of Molycorp ore...... 34

Table 3.1. Microflotation Reagents ...... 44

Table 3.2. WHIMS DOE Parameters ...... 46

Table 3.3. Falcon DOE Parameters ...... 47

Table 3.4. Bench Flotation Timetable ...... 48

Table 4.1. P80 of Ground Ore Samples ...... 50

Table 4.2. Modal Mineral Content of Major Components of Bastnaesite Ore. REE-bearing minerals are in bold...... 51

Table 4.3. Elemental Composition of Bastnaesite Ore ...... 51

Table 4.4. Mass and Bastnaesite Distributions. The mass percentage for each size fraction is given in plain text and the bastnaesite distribution in bold...... 53

Table 4.5. REE mineral and calcite liberation as cumulative mass recovery. Bolded values represent the combined total of REE minerals (bastnaesite, parisite, monazite, allanite)...... 53

Table 4.6. Calcium, barium, and TREE grade and recovery of WHIMS concentrate products...... 61

Table 4.7. Stream Concentrations ...... 64

Table 4.8. Shaking table mass balance for two similar 500 g tests, 2 and 3...... 65

Table 4.9. Mass balance for thrice-tabled concentrate. Products are the final concentrate and combined total middlings and tailings...... 67

Table 4.10. Signal-to-Noise Ratio for Falcon DOE Responses ...... 68

Table 4.11. ANOVA Table for the TREE Grade Model...... 68

Table 4.12. Grades, recoveries, and concentration ratios of Falcon products...... 69

Table 4.13. Bench Flotation Results: TREE and Calcium Grade and Recovery ... 70

Table 5.1. Comparison of Plant Annual and Capital Expense Differences (calculated as gravity-flotation plant values less flotation plant values)...... 72

Table 5.2. Concentrator economic model parameters (Operation)...... 74

Table 5.3. Concentrator economic model parameters (Capital Costs)...... 75

Table A.1. Size fraction volume and weight percentages ...... 83

Table A.2. Composition (weight percent) of each size fraction as reported by MLA analysis...... 84

Table A.3. Composition (mass percent) of each size fraction as reported by QEMSCAN...... 88

Table A.4. Overall calculated mineral abundance of the ground Molycorp bastnaesite ore...... 89

Table A.5. Composition comparison between QEMSCAN, MLA, XRD, and XRF...... 89

Table B.1. Microtrac particle size distributions for grind test samples ...... 91

Table B.2. Modal mineral content of the bastnaesite ore (wt%)...... 92

Table B.3. REE cumulative mass recovery to each liberation class ...... 97

Table C.1. Experimental data for microflotation tests ...... 98

Table C.2. Experimental data for WHIMS tests ...... 98

Table C.3. Experimental data for bench flotation tests ...... 98

Table C.4. Experimental data for Heavy Liquid tests ...... 99

Table C.5. Experimental data for Deister table tests ...... 99

Table C.6. Experimental data for Falcon tests ...... 100

Table D.1. Cash flow diagram for the flotation concentrator ...... 102

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Table D.2. Cash flow diagram for the gravity-flotation concentrator ...... 103

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CHAPTER 1: INTRODUCTION

In comparison to the broad spectrum of applications based on the rare earth elements, their supply shows little diversity. Their applications range from polishing media to hard drive magnets to water treatment additives to wind turbine motors. Prior to the 1950s the mineral providing the bulk of the rare earth supply was monazite; a phosphate mineral which was beneficiated primarily from placer deposits. There it could be separated from associated ilmenite, garnet, magnetite, quartz, rutile, and xenotime using a simple physical separation scheme utilizing differences in specific gravity, magnetism, and conductivity. For years this was the source material for cerium, lanthanum, neodymium, and thorium.[1]

In the 1950s the Mountain Pass Mine was developed by the Molybdenum Corporation of America, now Molycorp Inc. This deposit in southern California is the world’s richest source of bastnaesite, a fluorocarbonate containing cerium and lanthanum, as well as the heavy rare earths neodymium and praseodymium. Separation is much more difficult than with monazite from beach sands. This deposit contains bastnaesite and monazite as the primary valuable minerals (10% of the ore) with calcite and other carbonates (60%), barite (20%), and quartz and other minerals (10%).

Beneficiation of the rare earths from the bastnaesite ore at Mountain Pass involves concentration by flotation followed by roasting, leaching with hydrochloric acid, and solvent extraction. [1] Traditionally, Molycorp employed a crush, grind, float system to produce a 63% rare earth oxide (REO) concentrate, shown in Figure 1.1. The flotation conditioning involves alternating additions of steam, soda ash, lignin sulfonate, and tall oil, shown stepwise in Table 1.1. Flotation occurs at temperatures near 82 °C in several roughing, cleaning, and scavenging stages (Figure 1.2). The rougher produces a 30% REO concentrate which is further upgraded to 63% REO through cleaning. This is achieved at a 65-70% recovery.

Figure 1.1. Beneficiation Flowsheet at Mountain Pass (as of 1991) [2]

Table 1.1. Flotation Conditioning Steps at Mountain Pass. [3]

Amount Added Step Reagent(s) Added (kg/ton) Soda ash 2.5 - 3.3 Sodium fluosilicate 0.4 1 Steam (Bubbled through pulp) 2 Steam (Bubbled) Ammonium Lignin 2.5 - 3.3 3 Sulfonate Steam (Bubbled) 4 Steam (Bubbled) Distilled Tall Oil 0.3 5 Steam (Bubbled) 6 Steam (Bubbled)

Figure 1.2. Molycorp REO Flotation Circuit

China displaced the United States as the dominant supplier of rare earths in the mid-1980s due to production from the Bayan Obo mine in Inner Mongolia, and now is responsible for more than 90% of the world’s rare earths supply. Little effort has been put forth toward seeking a domestic source of the elements due to the cheap price of rare earths exported from China. Mountain Pass had been unable to compete with Chinese producers; it halted production in 2002 and resumed stockpile processing in 2008.

Hendrick compiled a report on U.S. rare earth commercial activity in 2008. No rare earths were mined, but stockpiled concentrate was processed at the Mountain Pass Mine. Bastnaesite concentrate and monazite prices, respectively compiled from USGS sources and U.S. import values, were $8,000/ton and $480/ton. [4] Again straining the supply in the US was the 2010 decision by the Chinese government to limit the volume of rare earth exports. This sanction caused a drastic spike in the price of the rare earth elements (shown in Figure 1.3) and brought questions to light of their availability, considering their importance in strategic applications.

Figure 1.3. REO Export Prices from China. Data retrieved from www.metal-pages.com

In 2011, the US Department of Energy released its Critical Materials Strategy; identifying rare earth materials as “critical” and in need of a more reliable domestic

4 supply. From this, the Critical Materials Institute was created with Molycorp as a partner in order to investigate their complex beneficiation problem. [5] Gleason’s 2011 article highlighted Molycorp’s strategic plan to resume production of rare earth minerals at Mountain Pass up to a goal of 40,000 metric tons/year by 2013. It also discussed some of the policy proposals laid out to accelerate the United States’ rare earth metals production. [6]

As of February 2015, Molycorp struggled to upgrade plant operations and compete with low-cost rare earth products from China. Rare earth oxide equivalent production by Molycorp for 2014 was 4,785 mt, which was an increase from 3,473 mt in 2013. [7] Rare earth prices had fallen dramatically from their highs in 2011. Due to this, Molycorp was forced to suspend operations in October of 2015 and operate only for care and maintenance of the operation. [8]

The general forecast is that an increase in rare earth production is needed to meet growing worldwide demand, and it will not be met from Chinese supply alone. More than 90% of the world’s rare earths come from China, but that comes from only 25% of the world’s reserves, accounting for recent estimates containing Canadian and Australian reserves.

Beneficiation of rare earths was recently summarized by Jordens, focusing on bastnaesite at Mountain Pass and Bayan Obo and monazite beach sands elsewhere in the world. Extensive Chinese knowledge in rare earth element (REE) flotation abounds in papers but many of them lack scientific depth and/or accuracy. Gravity concentration of rare earth minerals has been historically difficult due to the influence (and loss) of rare-earth-containing fines and the similar specific gravity of barite. [9] Limited commercial success has been seen with gravity separation of bastnaesite. The Sichuan Mianning REE Ore deposit uses shaking tables to separate minerals from pre-classified feeds. The mineralogy there is a carbonate containing barite, fluorite, and iron- and manganese-containing minerals with a 3.7% REO grade – most of which is bastnaesite – in coarse (>1mm) and fine powder (80% -325 mesh Tyler ) sizes. An all-gravity operation classifies and concentrates a 62% passing 200 mesh feed to achieve grades of 30%, 50%, and 60% with an overall recovery of 75%. Gravity and flotation are also

5 combined (Figure 1.4) to concentrate the feed and produce an overall 30% grade gravity concentrate at 74.5% REO recovery. The concentrate is then reground to 70% - 200 mesh and floated with hydroxamic acid, phthalate, sodium carbonate, and sodium silicate to produce a 50-60% REO grade concentrate with a rare earth recovery of 50- 60%. [10] Lab scale beneficiation examples are more abundant. A Mozley Multi-Gravity Separator was used to separate a Turkish bastnaesite ore, producing a 35.5% REO preconcentrate at 48% recovery. [11] Egyptian beach monazite was concentrated using electrostatic, magnetic, and gravitational methods. [12] The beach sands were screened to pass 1mm and deslimed then concentrated successively with a shaking table. The concentrate was subjected to low intensity magnetic separation and the nonmagnetic fraction beneficiated with a shaking table. The shaking table concentrate was dried and processed using high tension electrostatic separation, magnetic separation, and a shaking table again to produce a crude 85% monazite concentrate. More electrostatic and magnetic separation produces a 97% monazite concentrate (Figure 1.5). Humphrey spirals were used to concentrate an Iranian monazite ore. [13] Optimum results were found with an intermediate feed size, high feed rate (1.5 L/s), and low solids density (15%), with the latter parameter showing a less-significant effect. The best total rare earth elements (TREE) grade was reported at 6050x10-6 percent with a 57.06% recovery after gravity separation and leaching.

A developing site, Bear Lodge, owned by Rare Element Resources (RER) has achieved success with purely physical concentration. The deposit is a carbonatite, containing rare earths mostly as bastnaesite, with three regions of mineralization: oxide, sulfide, and transition. Drill core test work showed that the oxide core was successfully treated by scrubbing and sizing. The rare earth oxides were found to concentrate in the passing 500 mesh size. Results of scrubbing tests on the ore are shown in Table 1.2, where it can be seen that a 10-minute scrub yielded the maximized results in relation to recovery, weight rejected, and scrub time. [14]

Figure 1.4. Sichuan Mianning REE Deposit Flowsheet. [10]

Table 1.2. Characteristics of RER scrubbing tests. [14]

Scrub Scrub Scrub Scrub Scrub Characteristic Washed 5 min 10 min 15 min 20 min 60 min Weight % retained @ 52 50 40 37 48 52 90% recovery

REO recovery % @ 50% 82 89 91 92.5 88 87 passing

Assay grade @ 90% 13.5 14 16 13.5 17.5 13 recovery

Mesh size @ 90% 100 35 35 35 48 150 recovery

Head grade % TREO 9.11 7.82 7.15 6.21 9.27 8.62

% < 500 mesh 16.55 14.82 14.42 14.64 22.05 19.47

< 500 mesh 19.06 19.64 20.06 20.12 20.39 20.51 TREO grade

Figure 1.5. Egyptian Beach Sand Beneficiation Flowsheet. The first percentage given relates to the original weight and the second represents the assay of monazite. [12]

CHAPTER 2: LITERATURE REVIEW

Beneficiation of rare earth minerals has been reviewed previously. [15], [16] Special attention is given to the minerals (bastnaesite, monazite, xenotime, calcite, and barite), the collectors (oleic and hydroxamic acids), and the modifiers (soda ash, lignin sulfonate, metal salts, sodium silicate) used in this study. Methods of gravity concentration have also been surveyed. Prior to the discussion of reagents and their effect in flotation systems, a review of surface chemistry and flotation phenomena, as well as common analysis techniques, will be presented.

2.1 Flotation Surface Chemistry and Analysis Techniques

Flotation is a mineral processing technique that is widely used to concentrate an ore before recovery by pyrometallurgical or hydrometallurgical routes. It involves bubbling air through a tank of ground and crushed mineral pulp. Reagents are specifically chosen so as to cause selective adsorption of the desired mineral to the air bubbles. The mineral then rises through the pulp attached to the bubble where it sticks to the other bubbles in the froth. The froth is skimmed off or collected via overflow, and contains the upgraded ore. Often several steps are needed after concentration in the initial cell (“roughing”) to recover value from the concentrate (“cleaning”) or tails (“scavenging”). Several phenomena are at play, including the hydrophobicity of the minerals, the electrical potential of the minerals and solution, dissolution of mineral species, exposed surface area, and adsorption of reagents to mineral surfaces; many of which are affected by the temperature. Laboratory testing of flotation systems occurs in specially designed cells originally developed by A.F. Hallimond.

Fuerstenau collected data on contact angles, adsorption density, zeta potential, and flotation rate for a quartz-dodecylamine system over the full pH range. By plotting them together on sensible scales, their correlation was shown. That work refuted the claim that solid-liquid interface phenomena (adsorption density, zeta potential) cannot be connected to solid-liquid-gas interface phenomena (contact angle, flotation rate). [17] Hence many of these techniques can be used to predict overall flotation behavior. The

9 benefits and shortcomings of some of the industry’s common floatability tests have been previously detailed. There is no one industry standard, because all of the tests seem to have some sort of bias, whether it is operator expertise, non-industrial chemistry, unrealistic conditions, intensive time requirements, or some combination thereof. [18]

2.1.1 Hydrophobicity and Contact Angle Whether or not water attaches to a surface (of a mineral in this case) is qualified by its hydrophobicity, it’s “fear of water”. Water will attach to and wet a hydrophilic surface, but not attach to a hydrophobic substance. Sulfur and graphite are very hydrophobic, while calcite, quartz, and gypsum are hydrophilic materials. The source of this attraction is in the interfacial energies of the solid/air, solid/liquid, and liquid/air interfaces, related by Young’s Equation, represented visually in Figure 2.1:

Figure 2.1. Visual Representation of Young's Equation, showing the contact angle of a hydrophobic (left) and hydrophilic (right) surface. [19]

The contact angle can be experimentally determined using the sessile drop technique, which involves taking high-framerate consecutive images of a water drop as it is placed on a mineral surface. The geometry used to determine the contact angle is set up Figure 2.2. At the moment it hits the mineral surface, the water droplet forms a dome with a base diameter of d. If this dome were to be extended below the surface to form a sphere, that sphere’s radius would be R. The height of the dome above the mineral surface is given as a, and b is the difference in length between the sphere’s radius and the dome’s height (R minus a). The contact angle is then found at the tangent point T on the mineral surface using the triangle OBT.

O T

Figure 2.2. Illustration of the geometry used to determine the contact angle of water on carbon. [19]

Studies on bastnaesite, monazite, calcite and barite have shown that, like most minerals, they have hydrophilic surfaces. Flotation is a matter of selectively making mineral surfaces hydrophobic so they will become aerophilic, bind to the air bubbles, and float.

2.1.2 Hydrolysis Reactions When an ionic compound is placed in water, it will dissolve until equilibrium is reached. When sodium chloride does this, it leaves effectively no solid behind as the sodium and chlorine ions diffuse away from one another surrounded by shells of protons or electron pairs from water molecules. The result is a neutral solution. Some compounds, such as weak acids, equilibrate with a large concentration of the neutral species. The family of weak acids includes fatty acid collectors, like oleic acid. As it dissolves, a proton is taken away from the neutral acid and what remains is a charged molecule that can adsorb to an oppositely-charged mineral surface. Oleic and alkyl- hydroxamic acids can form salts with sodium and potassium to become sodium oleate and potassium octyl hydroxamate (when the alkyl chain is an octyl group). Collector structures and features will be detailed later. Another hydrolysis phenomenon takes place at the mineral surface, where exposed ions can attract charge-balancing H+ or OH- ions. The hydrolysis of cerium is particularly relevant in rare earth flotation: the

11 dissolution, hydrolysis, and readsorption of cerium is the primary attachment mechanism for hydroxamate collectors. [20]

The solubility product for the dissociation reaction HOl = H+ + Ol- at 20°C is 10- 10.9. Calcium and Barium oleate solubility products are 10-12.4 and 10-11.8. [21] Solubility products for hydroxamic acid and calcium and cerium hydroxamates are 109.35, 102.4, and 105.45. The equilibrium constant for the hydroxylation of trivalent cerium ions, 3+ - -20 Ce(OH)3 = Ce + 3OH , is 1.5x10 . [22]

The aqueous equilibria of cerium species is given in Figure 2.3. As the pH is increased, the species in equilibrium become less-positively charged. At a pH of around

10, the dominant species is the uncharged Ce(OH)3. At lower pH, the charged species 3+ 2+ + Ce , Ce(OH) , and Ce(OH)2 are dominant.

Figure 2.3. Aqueous Equilibria of cerium species. [22]

2.1.3 Zeta Potential and Point of Zero Charge For salt-type minerals, unequal dissolution can occur because the ionic components of the salt have different sizes and charges, which do not enter solution at the same rate. Because of this mineral surfaces are charged in solution. That surface attracts oppositely-charged ions from the bulk solution and creates an electrical potential. The ions manipulating the surface charge are called the potential-determining + - 2- 2- ions, and can consist of H , OH , CO3 , SO4 , and other ions (particularly those dissolving from the mineral). The Stern plane, illustrated in Figure 2.4, is the name given to the plane at which bound counter ions cannot come closer to the surface. The shear plane is the plane at which ions are capable of motion in the solution when forced. The potential at the Stern plane cannot be determined experimentally, but the potential at the shear plane can be, and is known as the zeta potential. At a specific pH this potential becomes zero. That pH is referred to as the PZC, the point of zero charge, for that mineral.

Figure 2.4. Schematic of electrical double layer (from Somasundaran 1975). [23]

The PZC is used as a marker for which kind of surfactant can be used with a mineral. At pH higher than the PZC, a surface appears positively charged (and the solution negative) as positive counter ions bind to the mineral surface, but the reverse is true below it. For example, oleate chemically adsorbs onto the surface of calcite above its point of zero charge, where the oleate ions bind with the calcium ions. Although the PZC is a critically important characteristic of a mineral, it often varies from sample to sample based on composition, crystallinity, and which plane of the mineral lattice is exposed (so which ions are free to dissociate). Table 2.1 demonstrates this variance for bastnaesite, Table 2.2 for monazite, and Table 2.3 for barite, calcite, and other semisoluble salt-type minerals.

Table 2.1. Bastnaesite Points of Zero Charge (PZC). [24]

Table 2.2. Monazite Points of Zero Charge (PZC). [24]

Figure 2.5. Zeta Potential of bastnaesite and monazite. (Luo Jiake 1984, from [16])

Table 2.3. Points of Zero Charge (PZC) of alkaline-earth semisoluble salt minerals. [24]

Figure 2.6. Zeta potential of barite, bastnaesite, and calcite in pure water. (Smith 1986, from [16]).

It has been determined that in the pH window for flotation, calcite is positively charged, bastnaesite is negatively charged, and the barite surface is undergoing transformation to a carbonate surface. [22]

Reagents are responsible for changes in the PZC of minerals during flotation by adhering to their surfaces and thus changing their charge. Bastnaesite, barite, and calcite surface chemistry has been analyzed in response to changes in soda ash concentration (used to modify the pH). The carbonate ion had the most pronounced effect on the bastnaesite zeta potential, and the least on the calcite potential. The barite zeta potential changed dramatically from positive to negative at 8x10-4 M carbonate due to the formation of barium carbonate on the mineral surface. [25] Initially, the zeta potential of the bastnaesite is more negative with respect to barite, but the barite becomes more negative after an addition of 1x10-4 kmol/m3 ammonium lignin sulfonate, because of the stronger adsorption onto the barite surface. [26]

Cheng’s PZC of monazite was reported as a pH of 5.3. This, combined with the negative zeta potential at high pH, leads to the conclusion of oleate ions being chemisorbed onto the surface. [27] The monazite zeta potential curves are shifted to the left (the PZC becomes lower) and made steeper in the presence of hydroxamate collectors. [28] Cheng also discussed computer modeling of CePO4 and YPO4 as the

16 basis for the wide range of reported PZC values (pH 3-7) for monazite and xenotime, respectively. An alternative cause for variation was given: impurities from other ore bodies and crystallinity, which determines which ions are exposed. [29]

2.1.4 Adsorption Density Surfactants can bind to mineral surfaces in different configurations. The mechanisms for surface attachment can be classified as low- or high-energy processes. Physical adsorption involves van der Waals bonding and hydrogen bonding, while the higher energy chemical adsorption relies upon covalent bonding. Molecules can attach themselves horizontally, leading to lower adsorption densities; or in vertical configurations, leading to higher adsorption densities. Multiple layers can form as the molecules match hydrophobic ends or hydrogen bond from one chain to another. As many as six hydroxamate layers have been reported to form at the interface. [22] Adsorption densities on barite (as a horizontal monolayer) and on calcite (as a horizontal layer, then a vertical layer) are much lower than on bastnaesite. Calcite exhibits the curious characteristic of a linear increase in adsorption perhaps, but unconfirmed, as the result of calcium hydroxamate precipitation.

Figure 2.7. Oleate adsorption onto bastnaesite as a function of pH with and without pre- boiling. (Smith 1986 from [16]

Figure 2.8. Adsorption isotherm of oleate on calcite at pH 9.3. [21]

Figure 2.9. Adsorption Density of hydroxamate on barite, calcite, and bastnaesite at 21°C and pH 9.3. [22]

Adsorptions of hydroxamate were found to be endothermic. [22] The free energy of adsorption was estimated as -26, -28, and -57 kJ/mol for barite, calcite, and

18 bastnaesite. Bastnaesite’s trivalent state (as opposed to the divalent state of the alkaline-earths) is proposed as a factor in the increased adsorption. [20]

FTIR has been used to distinguish whether physical adsorption or chemisorption occurs between monazite and bastnaesite and sodium oleate and potassium octyl hydroxamate. Sodium oleate physically adsorbs onto bastnaesite and monazite at pH 9 and pH 3 and 8, respectively. Potassium octyl hydroxamate adsorbs chemically onto both minerals at pH 9.3 and 9. FTIR was unable to distinguish whether or not chemisorption at pH 8 occurs for monazite and sodium oleate. [30]

2.1.5 Hallimond Tube Flotation Fuerstenau modified the design of the original Hallimond Tube to a version similar to that used in this study. [31] It is one of the most common devices used for microflotation tests, although due to its simplified design, it is not considered representative of industrial flotation. [18] Several reasons for this are incomplete chemistry (a frother is not necessary, difficulty in generating reliable grade-recovery curves, and unrealistic flow conditions.

Although operation of a Hallimond tube is simple, the parameters are not standardized. Gas composition and flow rate, stirring speed, specific water volume and temperature, reagent additions, and flotation time are all determined by the researcher. The tube is filled with a very dilute slurry (one gram of mineral per 100 mL of water), and gas is bubbled through a frit at the bottom. A magnetic stir bar disperses the bubbles, which attach to particles in the slurry. As the bubbles rise to the top of the water, they break, and the mineral falls into the concentrate stem.

2.1.6 Temperature Effects The literature shows that for all three minerals, collection increases with increased temperature (Figure 2.11). Bastnaesite shows a more pronounced effect. This is the impetus for Molycorp’s steam conditioning: increasing selectivity. The endothermic nature of adsorption revealed that it is a chemical adsorption process and temperature was proposed as a driver of the increased adsorption. [32] Flotation with elevated conditioning temperatures has shown (Figure 2.12) that between hydroxamate 19 and fatty acids, the selectivity for bastnaesite increases with temperature, but more so for hydroxamate. [22], [32]

Figure 2.10. Schematic Drawing of a Modified Hallimond Tube. [31]

Figure 2.11. Effect of pulp temperature on oleate flotation.

Figure 2.12. The flotation recoveries of Mountain Pass ore as a function of conditioning temperature for fatty acids (left) and hydroxamate (right) collectors. [22]

2.2 Minerals

Hanna and Somasundaran discussed many aspects of the flotation behavior of salt-type minerals, including mineral crystal structure, dissociation, and surface hydroxylation. [21] Surface charging and zeta potential were discussed with regard to calcite and other minerals. Interactions of several depressants, including metal salts and sodium silicate, were included. Surface characteristics of rare earth semisoluble salt minerals were discussed as well. [24] Electrokinetic properties, surface wetting, and collector adsorption behavior are detailed in later technical sections.

2.2.1 Bastnaesite Bastnaesite is a rare-earth fluorocarbonate mineral with a formula of

(Ce,La)FCO3. Cerium occurs in the mineral more often than lanthanum. It has been found in numerous deposits around the world; the two most notable of which are Bayan Obo in China and Mountain Pass in California. [1] It has a specific gravity of 5.0.

As the predominant supplying mineral of rare earth elements in the world its beneficiation has been extensively studied and reviewed. [3], [9], [22] Pradip performed electrokinetic, Hallimond tube and Denver cell flotation, adsorption, and x-ray studies in his thesis. The surface chemistry of bastnaesite, barite, and calcite has been analyzed. Flotation experiments have been conducted with fatty acids and hydroxamic acids, along with additions of inorganic salts, organic ions and molecules, soda ash, and lignin sulfonate.

2.2.2 Monazite The initial source of rare earth elements, monazite, is a rare-earth phosphate

(La,Ce)PO4. It was originally beneficiated as a nuclear reactor material from placer deposits using gravity, magnetic, and electronic separation techniques. When bastnaesite containing much lower amounts of thorium was discovered, monazite fell out of fashion. It is found in deposits containing bastnaesite, and thus necessitates a separation step for those ores. Monazite’s specific gravity varies from 5.0 to 5.4.

Pavez and Peres tested species of monazite, zircon, and rutile using three different collectors (sodium oleate, potassium octyl hydroxamate, and a commercial hydroxamate) and a depressant (sodium metasilicate). [28]

2.2.3 Calcite and Barite

Calcite is a carbonate mineral of calcium, CaCO3. It is a semisoluble salt-type mineral, and has been the subject of numerous investigations. [33], [34], [21], [35], [25] It is present in Mountain Pass ore as one of the primary gangue constituents. It has a specific gravity of 2.7.

The other major gangue mineral is the sulfate barite, BaSO4. It has also been studied, often in attempt to depress it from bastnaesite. [21], [22], [25], [26] Barite has a specific gravity of 4.5.

2.3 Reagents

Industrial flotation requires the use of surface-modifying chemicals in order to efficiently separate the desired minerals from the gangue. Collectors are those used to target the desired minerals, while modifiers and depressants are used either to promote flotation of the desired mineral or to inhibit gangue flotation.

2.3.1 Collectors Mineral flotation relies on collectors to attach to desired mineral surfaces and bubbles. In sulfide flotation, the dominant collector type are xanthates. In rare earth flotation, the traditional molecules used are fatty acids, while hydroxamates are a promising group undergoing laboratory study, with limited industrial application.

2.3.1.1 Fatty Acids

A fatty acid is an organic acid consisting of a chain of singly or doubly bonded carbon atoms and a functional group capable of donating a proton. As fatty acid chain length increases from 8 to 12 carbons, collector concentration required for flotation decreases (Figure 2.13). [34] As shown in Figure 2.14, at a concentration of 10-3 mol/L fatty acid, chain lengths of 11 and 12 carbons float well from pH 6-12.5. Smaller chains show minimal recovery above pH 10. The proposed mechanism for the collector - adsorption to surface calcium and carbonate is by the reaction CaCO3 + 2(RCOO )  2- Ca(RCOO)2 + CO3 .

The structure of oleic acid is shown in Figure 2.15. It is a monounsaturated 18- carbon chain ending in a carboxyl group (a double-bonded oxygen and a hydroxyl group are bonded to the end carbon). Flotation of barite, calcite, and fluorite was studied using oleic acid and sodium oleate (the sodium salt formed by oleic acid and sodium ions). Electrokinetic, Hallimond tube, and abstraction tests suggest that a layer of calcium oleate forms around the minerals, which prevents further dissolution to equilibrium. Due to their similar characteristics, selectivity between the three minerals was proposed as unlikely to be obtained. [35]

Figure 2.13. Calcite recovery as a function of fatty acid addition for various fatty acids at pH 9.7. [34]

Figure 2.14. Calcite recovery as a function of pH for various fatty acids at a constant collector addition of 10-3 mol/liter. [34]

Figure 2.15. Structure of oleic acid

Figure 2.16. Bastnaesite recovery by sodium oleate as a function of pH. [26]

Recovery of bastnaesite with and without additions of a depressant is shown in Figure 2.16. The maximum is in the slightly-alkaline pH region. Bastnaesite, barite, and calcite all float best around pH of 9.5 but bastnaesite flotation requires a lower concentration of sodium oleate (3x10-4). By combining that fact with the calcite minimum around pH 8.5, conditions for effective separation have been determined, although plant practice dictates the use of depressants. [22]

The maximum floatability of monazite was shown to coincide with the maximum concentration of Ce(OH)2+ and La(OH)2+ (from thermodynamic modeling), the greatest particle-bubble adhesion, and a pH of 8.5 – 9. [27]

Sodium oleate floatability experienced a minimum between pH of 4-5. Maxima exist on either side for monazite (3 and 7), zircon (3, 6-8), and rutile (3, 7-8). Hydroxamate floatability occurs between 3 and 7, 2.5- 9, and 3-8, for monazite, zircon, and rutile. [28] Shown in Figure 2.17, the percentage of calcite floated was 20% at 3x10- 6 M, 80% at 6x10-6M, and 95% at 10-5 M oleate concentration. [21]

Figure 2.17. Oleate concentration required to float salt-type minerals.[21]

2.3.1.2 Hydroxamates

Hydroxamate collectors are chelating molecules that contain several active sites to which an ion can bond and a sufficiently long chain to provide the necessary hydrophobicity. That chain (represented by R in Figure 23) can be an alkyl group with 7- 14 carbons, a naphthalene ring, or other organic group. The number of active sites varies from specific molecule to molecule, but they all function by forming coordinating bonds with metal ions. [36] The specific mechanism of attachment can also vary but an example is given in Figure 2.18, where the hydroxyl group is deprotonated and the two oxygens coordinate to bind the metal ion. Different layer formations (either monolayer or multilayer) are the result of horizontally versus vertically oriented molecules. These layers form due to chemisorption with surface cations: cations form hydroxy complexes, readsorb, then bond with the hydroxamate. [20]

Figure 2.18. A mechanism of hydroxamate adsorption to an ion on a mineral surface. [20]

For his thesis Pradip detailed the preparation of potassium octyl hydroxamate (structure, Figure 2.19) and characterized the reagent used as being composed of 50% potassium octyl hydroxamate and 50% octyl hydroxamic acid. [22]

Figure 2.19. The structure of potassium octyl hydroxamate. [22]

Modified hydroxamic acid (R-OHCOONH2, where R is a naphthalene ring) chelates the cerium on the bastnaesite surface, yielding best flotation between pH 8 – 9.5 (Figure 2.20). [37] The primary factor limiting the use of these more selective reagents has been their high cost. [36]

Figure 2.20. Modified hydroxamic acid chelating surface cerium(III). [37]

Figure 2.21 demonstrates the floatability of bastnaesite with hydroxamates between pH 6 and pH 9. Denver cell flotation based on the Hallimond tube results showed the increased selectivity (at both room temperature and elevated temperature) of the hydroxamate as compared to the fatty acid. [22]

Figure 2.21. Hallimond tube flotation results on bastnaesite with K-octyl hydroxamate as collector. [22]

Figure 2.22. Bastnaesite ore recovery (as total weight, including gangue, recovered) by K-octyl hydroxamate as a function of pH. [22]

2.3.2 Modifiers and Depressants Collectors alone are rarely sufficient to achieve desired selectivity in flotation. A variety of compounds exist to modify the surface and/or solution chemistry of a mineral flotation system. The pH must be regulated to create the proper electrokinetic environment for the collectors to attach to the minerals. As bastnaesite flotation occurs in alkaline environments, a base (hydroxides and soda ash are common) must be added to raise the pH. Other chemicals are used to inhibit flotation of a specific mineral; these are depressants.

2.3.2.1 Soda Ash

Soda ash (Na2CO3) is used to regulate pH. The carbonate ion is a potential- determining ion for bastnaesite and calcite. In the flotation window of 8 < pH < 10, HCO- - 2- 3 and CO3 are present in solution. At high enough concentrations, the surface of barite is converted to a barium carbonate surface. Excess soda ash can lead to depression of bastnaesite. [22]

2.3.2.2 Lignin Sulfonate

Lignin sulfonate is a complex compound derived from the sulfonation of lignin, a cellulose binder, in wood pulp. It is known to flotation as a barite depressant. [21], [22], [26] Its selectivity for barite at high pH is likely due to the highly positive surface of the barite compared to that of bastnaesite and calcite. It has also been proposed that the molecule fits better on the barium sulfate structure as compared to the carbonate structures of bastnaesite and calcite.

Bastnaesite recovery as a function of ammonium lignin sulfonate concentration is given in Figure 2.23. Barite flotation with 1x10-5 kmol/m3 sodium oleate is virtually eliminated by 1x10-5 kmol/m3 ammonium lignin sulfonate.

2.3.2.3 Metal Salts

Inorganic salts can increase flotation by binding and providing new adsorption sites or decrease it by competing with collector ions. Salts composed of potential- 2- determining ions can manipulate the charge of minerals in the system, such as SO4 for

2- barite and CO3 for calcite and bastnaesite. [22] Gaudin studied calcite flotation with various metal salts, showing that many of them depress calcite flotation by causing precipitation of the oleic and undecylic acid collector. The order of their effectiveness was found to be Cu(NO3)2•3H2O > Na2SiO3 > CuSO4•5H2O > FeSO4•7H2O. [33]

Figure 2.23: Bastnaesite recovery as a function of ammonium lignin sulfonate concentration. [26]

Figure 2.24. Recovery of calcite as a function of increased concentration of metal salts. [33]

2.3.2.4 Sodium Silicate

Sodium silicate is well known as an inorganic modifier. In rare earth mineral flotation, sodium metasilicate has been found to be an effective zircon and rutile depressant while not affecting monazite. [28] It was shown to be an effective depressant of zircon and rutile (Figure 2.25), more so than sodium sulfide. A temperature increase slightly increased selectivity. [38] The increase in sodium silicate adsorption and decrease in oleate adsorption are shown to reduce flotation recovery in

Figure 2.26. It increases flotation of barite and calcite from quartz, but can depress calcite with respect to fluorite, especially when aluminum salts are used. [21] Figure 2.27 shows depression of barite with respect to rare earth minerals and fluorite.

Figure 2.25: Effect of sodium metasilicate for oleate (top) and hydroxamate (bottom) flotation. [38]

Figure 2.26. Adsorption and Floatability of sodium oleate and sodium silicate on calcite and fluorite as a function of sodium silicate concentration. [21]

Figure 2.27. Floatability of different minerals versus Na2SiO3. (Luo Jiake 1984, from [16] 32

2.4 Gravity Concentration

Often before flotation is considered, gravity separation is investigated as a means of concentrating an ore. In the case of liberated, similarly-sized beach sands or free gold deposits where the desired mineral is much denser than the undesired minerals, gravity can quickly sort the valuable material from the gangue. This approach relies on complete liberation – distinct particles of the valuable component must exist without contact with gangue particles. These separations generally show better results with a near-size deslimed feed. When a disparity exists between the specific gravities of the two components, gravity may be used to sort them.

The Concentration Criterion is used as a first estimate of the relative success of gravity separation. It compares the specific gravities of the heavy particles Dh, light particles Dl, and fluid (usually water) Df.

Concentration criteria greater than 2.5 indicate that gravity separation is viable, while those below 1.25 mean it is practically impossible. Values between those suggest that using the right equipment and a carefully controlled feed, a separation could be made. [39] Table 2.4 shows the concentration criterion for major components of the bastnaesite ore used in this study.

Not every type of gravity equipment is suitable for a given application, which is why so many exist. There are vibratory motion-based devices, such as jigs and shaking tables, centrifugal units like Falcon and Knelson concentrators, and other devices: spirals, multi-gravity separators, and heavy media separators. Characteristics of the equipment such as allowable feed size (Figure 2.28), throughput, water requirement, plant footprint, and power requirement dictate their applicability to a given separation. Following the Concentration Criterion calculation, a float sink analysis may be done to determine the efficiency of the separation as a function of specific gravity.

Table 2.4. Specific Gravity (SG) and Concentration Criterion (CC) for major components of Molycorp ore.

Mineral Formula SG CC

Monazite (Ce,La)PO4 5.2 2.47

Bastnaesite (Ce,La)FCO3 5.0 2.35

Barite BaSO4 4.5 2.06

Parisite Ca(Ce,La)2(CO3)3F2 4.4 2.00

Rutile/Anatase TiO2 4.1 1.81

Strontianite SrCO3 3.8 1.65

Ankerite Ca(Fe,Mg)(CO3)2 3.0 1.18

Chlorite (Mg,Al,Fe)12[(Si,Al)8O20](OH)18 2.9 1.12

Calcite CaCO3 2.7 1.00

Quartz SiO2 2.7 1.00 Feldspar (K,Na,Ca…)X(Al,Si)3O8 2.7 1.00

These tools, along with preliminary testwork, can be used to assess the type of equipment suitable to beneficiate an ore. In the case of rare earth ores, the recent development of centrifugal-type concentrators has pushed the boundary of allowable separation into the domain at which these minerals concentrate. As the minerals require liberation, they must be ground finely. Excessive fines can be problematic in flotation due to entrainment and agglomeration issues, although dispersants can be used to mediate that effect. The slimes can also reduce the sharpness of the separation, as seen in Figure 34. While potentially a problem for some types of equipment, spinning- bowl concentrators are able to process feeds as fine as tens of microns. They are joined by hydrocyclones, tilting frames, Mozley tables, and froth flotation as the only nonmagnetic unit operations, according to Figure 2.28, capable of handling a feed of that size. Wet tables reach into the top end of this range, along with many other operations. Two carbonatite operations similar to the Mountain Pass deposit that use physical beneficiation methods are Sichuan Mianning (shaking tables and flotation) [10] and Bear Lodge (scrubbing and sizing) [14].

Molycorp bastnaesite ore has seen prior attempts at shaking table concentration in a lab setting. [40] The bastnaesite ore was pulverized and split into four fractions: 20- 38 μm, 38-53 μm, 53-75 μm, and 75-106 μm then purified with a Frantz Isodynamic

Separator. That was then processed on a Mozley shaking table at 90 rpm with a 3.5” stroke and 3L/min of wash water. These methods yielded a relatively pure bastnaesite sample in the 38-53 μm and 53-75 μm fractions. Figure 35 shows the XRD patterns of those relatively pure samples. Outside of those size fractions, considerable amounts of calcite, barite, and quartz were observed.

Figure 2.28. Applicability of Beneficiation Equipment to a given feed size. [41]

Heavy liquid separation will be elucidated due to its importance in characterizing the gravity response of an ore. Shaking tables are a tested and proven method of separating an ore, with a slightly finer size range than jigs and spirals. Here, centrifugal concentrators were used to demonstrate their ability to effectively separate a feed of such fine size.

Figure 2.29. Effect of contaminating slimes and suspension viscosity on the cleaning of Roslyn coal. [42]

Figure 2.30. XRD Patterns of Molycorp Bastnaesite size fractions upgraded with a Frantz Isodynamic Separator and reference minerals. All diffraction peaks have been normalized to the maximum peak intensity for each pattern. [40]

2.4.1 Heavy Liquid Separation To quote Chris Mills, “The first step at the laboratory level should always be heavy liquid analysis of the ore to be fed to the gravity separation plant.” The information gained from such tests can dictate whether gravity separation will be easy or difficult and which types of equipment are available to make the separation. [41] While many fluids exist in the specific gravity range of 1.2 - 2.0, options available for gravity separation of ores are both more limited and more expensive. [42] Historically, potentially hazardous halogenated hydrocarbons have been used for this type of work. Sodium polytungstate, a newer, nontoxic reagent with a maximum s.g. of 3.1 (adjustable by means of the water-to-powder ratio) was used in float/sink analysis. [43]

Figure 2.31. Density of aqueous sodium polytungstate solution as a function of mass percent. [43]

2.4.2 Shaking Tables Shaking tables are rectangular-shaped tables with riffled decks across which a film of water flows. (Figure 2.32 and Figure 2.34) The mechanical drive imparts motion along the long axis of the table, perpendicular to the flow of the water. [44] The water carries the particles of the feed in slurry across the riffles in a fluid film. This causes the fine, high density particles to fall into beds behind the riffles as the coarse, low-density particles are carried in the quickly-moving film. (Figure 2.33) The action of the table is such that particles move with the bed towards the discharge end until the end of the

37 table stroke, at which point the table rapidly moves backwards and the particles’ momentum propels them still forward.

Figure 2.32. Deister shaking table deck. Dark circles represent low-s.g. particles. White circles represent high-s.g. particles. The diameter of the circles corresponds of the particle size. [45]

Figure 2.33. Fluid Motion over riffles on a Deister deck. [45]

The feed characteristics, feed rate, riffle pattern, and motion of the table should all be tailored to fit the desired application. Feed characteristics are generally set by the comminution circuit, but classifying, either with screens or cyclones, can influence the separation on the table. Riffle pattern is most easily controlled by changing decks: often a sands deck is used for coarse feeds and a slimes deck is used for fine feeds. The drive of the table can be manipulated in both stroke length and frequency. A longer stroke will require more water but moves heavies to the concentrate end more quickly. Deister decks are set up at an incline from the drive to discharge end to allow migration of heavies to the concentrate end and allow light particles to fall to the tails easier. The tilt from the dressing side to the tailings side is generally maintained to allow a wide spread of material at the concentrate end. While the tails-middlings cut point is dictated by collection bins around the table, the concentrate-middlings cut point is made by the operator at some point along the discharge end.

Figure 2.34. Wilfley shaking table

2.4.3 Knelson and Falcon Concentrators Development of centrifugal concentrators was pioneered by those searching to separate free gravity recoverable gold (GRG). In the 1970s, the Knelson Bowl (Figure 2.35) and later the Falcon Concentrator (Figure 2.37) were developed to use centrifugal force to amplify the force of gravity for the purpose of separating constituents of an ore.

A rotor spins the bowl, which throws the feed coming into the center against the walls of the bowl. Light and fine particles are carried out of the bowl with the tailings while heavy and coarse particles are collected and removed from the bowl. Industrially they can operate as batch or continuous units, with pores opening intermittently to collect concentrate.

Figure 2.35. Knelson Concentrator schematic. [46]

The Knelson Bowl is the more widely-accepted centrifugal unit. The difference between a Falcon and Knelson lies in the proprietary design of the bowl. Factors such as bowl height, wall angle, radius of the bowl sections, and size and placement of collection ports vary between the two brands. In particular, the Knelson utilizes concentrate collection all along the height of the bowl, while a Falcon collects the concentrate in several rings at the top of the bowl. Laplante’s Standardized GRG test used a Knelson to separate and classify feeds and determine the amount of GRG in

40 each size fraction, as well as the overall susceptibility to gravity recovery. [47] Exemplary curves for a poor, intermediate, and exceptional response are shown in Figure 2.36 as lines a, b, and c.

Figure 2.36. Cumulative gold retained as a function of particle size. [47]

Falcons have been shown to have dependence on slurry density only to a slight degree as the particles are quickly forced from the fluid and into the bed. There is a positive relationship between flow rate and recovery up to a plateau around 20 L/min. At low slurry feed rates and densities (10 L/min and 10% solids) the Falcon recovery suffered. [48] Behavior of slimes in a Falcon has been modeled previously and shown to correlate with industrial experimental data. [49]

Figure 2.37. Cross-Section of a Falcon Concentrator

CHAPTER 3: EXPERIMENTAL METHODS

Once the literature survey was completed, an experimental campaign was developed and carried out to determine the effectiveness of gravity separation before froth flotation.

3.1 Characterization and Mineralogy Procedures

Ore was provided in two lots by Molycorp Inc. The first was crushed in a jaw and roll crusher, then ground batch-wise in a laboratory ball mill until it was 100% passing 100 mesh. This ore was then blended and split in a Jones Riffle, and was used for the microflotation tests. The second was crushed in a roll crusher until it passed 12 mesh then was blended and split into representative samples in a Jones Riffle. Those samples were wet ground in a rod mill for the required length of time.

Samples of the ore were sent to the Center for Advanced Mineral and Metallurgical Processing (CAMP) at Montana Tech and to the Colorado School of Mines Geology Department for MLA and QEMSCAN analysis, respectively. Several samples, ground in a rod mill for specific lengths of time (zero, 10, 30, 60, and 90 minutes), were characterized by CAMP to determine the liberation behavior of the ore components as a function of grinding time. Elemental composition was determined by the automated mineralogy software (as part of MLA and QEMSCAN), x-ray fluorescence, and x-ray diffraction. The microflotation samples were analyzed with the Kroll Institute XRF, while all of the gravity and magnetic test samples were analyzed by Hazen Research, Inc. Size analysis was performed with a Microtrac Particle Size Analyzer and several Tyler sieves.

3.2 Microflotation

Sodium oleate, octanohydroxamic acid, soda ash, ammonium lignin sulfonate, sodium silicate, and copper (II) nitrate hemipentahydrate were procured as solids and dissolved in de-ionized water to create stock reagent solutions. Concentrations (listed in

Table 3.1) were chosen based on previous studies on single-mineral systems. Research purity nitrogen was chosen to minimize the gas species present in the system and was obtained from General Air.

Ore grind size, the use of soda ash for pH control, oleic acid as a collector, a sulfonate depressant, and flotation at elevated temperatures were based off of Mountain Pass plant practice (Figure 2.12 and Figure 2.16). [3], [22] Octyl hydroxamate was chosen based on reports of increased selectivity as compared to oleate (Figure 2.12). [20], [22], [28], [32], [36] Ammonium lignin sulfonate was used as a barite depressant (Figure 2.23). [26] Copper nitrate was chosen due to previous depression of calcite (Figure 2.24). [33] Sodium silicate was used to depress gangue, particularly from monazite (Figure 2.25). [38] Microflotation tests were performed in a Hallimond Tube (similar to that in Figure 2.10).

Table 3.1. Microflotation Reagents

mol/L mg/L Octano- Test Sodium hydroxamic Ammounium Copper (II) Sodium Number Test Name Oleate Acid Lignin Sulfonate Nitrate Silicate 1 NaOl 11 1.00E-05 2 NaOl 3 5.00E-05 3 NaOl 14 1.00E-04 4 H9 1.00E-04 5 H5 3.00E-04 6 H1 5.00E-04 7 NaOl 2 5.00E-05 5.00E-06 8 NaOl 13 5.00E-05 5.00E-06 9 NaOl 4 5.00E-05 1.00E-05 10 NaOl 7 5.00E-05 2.00E-05 11 H10 3.00E-04 5.00E-06 12 H6 3.00E-04 1.00E-05 13 H13 3.00E-04 2.00E-05 14 NaOl 8 5.00E-05 100 15 NaOl 5 5.00E-05 300 16 NaOl 10 5.00E-05 500 17 H11 3.00E-04 100 18 H7 3.00E-04 300 19 H14 3.00E-04 500 20 NaOl 12 5.00E-05 1.00E-07 21 NaOl 6 5.00E-05 5.00E-07 22 NaOl 9 5.00E-05 1.00E-06 23 H12 3.00E-04 1.00E-07 24 H8 3.00E-04 5.00E-07 25 H15 3.00E-04 1.00E-06

One-gram samples were taken in batches from a bag (approximately 15-20 grams per batch) and the grade of those batches was recorded. Conditioning was performed in two separate 150-ml beakers. This was done to limit the amount of slurry entering the collecting arm of the tube, therefore artificially inflating the recovery numbers. The volumes were selected based on the volume of the cell. At around 40 ml, solution began to pour into the collecting arm. As such, 25 ml of water was targeted for the slurry beaker (ore, water, and depressant), although this number varied slightly as more or less stock depressant solution was used to achieve the desired concentration. The solutions were heated on a hot plate as the slurry was stirred with a stir bar. After ten minutes (at a temperature of 80±10°C), the collector was added to the slurry and pH was adjusted to 9.0±0.1 using drops of soda ash solution to follow plant practice. After fifteen minutes, the slurry was removed from the hot plate and poured into the Hallimond tube. The remaining solution was then poured in.

Flotation lasted for two minutes at a flow rate of 60 cc/min nitrogen gas. The slurries were stirred with a stir bar to ensure mixing. The apparatus used for the tests is shown in Figure 3.1. After that time, the concentrate was gathered by removing the stopper from the collecting arm and flushing the froth from the tube. The remaining tails solution was collected as well. Both the concentrate and tails were filtered through Whatman 40 (8-μm pore size) filter paper and dried for 18 hours.

Figure 3.1. Modified Hallimond Tube used for microflotation tests.

3.3 Magnetic Separation

Fifty grams of ore ground for zero, 30, and 90 minutes were charged at 10% solids through the WHIMS at different field strengths, based on percentages of maximum amperage. The two-factor DOE matrix (designed with Stat Ease) is given in Table 3.2.

Table 3.2. WHIMS DOE Parameters

DOE Field Strength, Standard WHIMS Run P80, µm Gauss 5 1 762 5000 4 2 50 5000 6 3 762 10000 1 4 144 7500 3 5 50 10000 2 6 144 7500

3.4 Gravity Concentration

Sodium polytungstate heavy liquid medium was used to investigate the possibility of beneficiation by gravity separation. Five grams of ore (ground for zero, 30, and 90 minutes) were centrifuged in 10 mL of fluid of specific gravity ranging from 2.70 to 2.95. The floats were poured and skimmed off. The middling solution was poured off, and the fines were flushed. Each product (floats, middlings, and sinks) was washed and filtered several times and dried. The composition of the products was determined with x-ray fluorescence.

A Deister table was used to develop a qualitative response to gravity concentration. The process variables were adjusted based on visual observation. An initial 500 g feed was tabled, yielding a concentrate, middling, and tailing product. That concentrate was tabled again to represent a cleaning step. Two more 500 g batches were tabled under the same conditions. From one of those tests, the concentrate, middlings, and tailings were tabled again. A 300 g charge of 100 x 325 mesh ore was also processed to investigate the effect of a classified feed. Fluidization water was set at

46 the minimum level that created a film across the entire table. The ore was slurried and hand-fed from a small bucket into the feed box. Table tilt was adjusted as the feed spread across the table. The cut point for concentrate and middlings was made at the discretion of the operator based on the visual quality of the film. The table elevation, stroke length, and stroke frequency were kept constant throughout all tests.

Stat Ease Design-Expert 9 was used to generate a Design of Experiments matrix for work on the Falcon Concentrator. The factors chosen (Table 3.3) were G-Force (controlled by the frequency on the Variac controller attached to the Falcon motor), Feed Rate (controlled by opening the valve between one-half and one full turn open), and Feed P80 in microns (controlled by grinding for specific lengths of time). The slurry density was maintained at 10% solids for all tests, although that was approximated in later tests due to uncertain moisture content. The slurry was mixed in a 20-liter tank and fed through a valve to the Falcon.

Table 3.3. Falcon DOE Parameters

DOE Falcon G-Force, Grind Time, Feed Rate, Standard Run G's minutes kg/hr Opt 1 100 90 30 Opt 2 100 90 30 Opt 3 100 90 30 1 1 100 90 30 6 2 250 30 30 4 3 250 90 60 3 4 100 90 60 9 5 175 60 45 8 6 250 30 60 11 7 175 60 45 2 8 250 90 30 5 9 100 30 30 7 10 100 30 60 10 11 175 60 45

3.5 Bench Flotation

Bench Flotation tests were performed to compare the result of gravity preconcentration to a circuit with no such step. A benchmark test using ore ground to a

P80 of 50 μm was carried out using lignin sulfonate and hydroxamic acid. Each test was performed according to the timetable of Table 3.4. The slurry was heated at 33% solids to approximately 85°C, and transferred to the flotation cell for conditioning. Ammonium lignin sulfonate and octanohydroxamic acid were added and the motor turned on once the slurry was in the heated cell. After fourteen minutes one drop of methyl isobutyl carbinol was added to encourage frothing. At fifteen minutes, the air valve was opened and flotation began. Flotation continued for ten minutes, with scraping of the froth occurring every thirty seconds. For all tests, the air flowrate was 280 ml/min and the motor stirred at 900 RPM. Following the tests, samples were filtered, dried, weighed, and analyzed by XRF. Those conditions were repeated for a flotation test of the gravity concentrate.

Table 3.4. Bench Flotation Timetable

Time Event Start Beginning of conditioning: turn motor on; add slurry, collector, and depressant; adjust pH 14 minutes Addition of frother, adjust pH 15 minutes Beginning of flotation: Turn air on 16 minutes Scrape froth, (repeat every 30 seconds) 25 minutes End of flotation: turn air and motor off

CHAPTER 4: RESULTS AND DISCUSSION

The data received from the experiments was collected and analyzed, and is presented below along with discussion of its interpretation.

4.1 Characterization and Mineralogy

Mineralogy results were received for the two lots of ore submitted for characterization by MLA and QEMSCAN. The initial lot was used for the microflotation and shaking table concentration and the second was used for the Falcon concentration, magnetic separation, and comparative flotation tests. The results for the second lot are presented in this section and those for the second lot can be found in FIRST LOT MINERALOG.

The particle size distribution and P80 for each grind time is are given in Figure 4.1 and Table 4.1. These distributions were generated by grinding batches of ore for the specific time (zero, 10, 30, 60, and 90 minutes). Subsequent test batches were ground for similar lengths of time based on the intended distribution and size at which point

80% of the material passes (P80).

Figure 4.1. Particle size distributions for ground ore samples

Table 4.1. P80 of Ground Ore Samples

Grind Time (minutes) P80 (microns) 0 762 10 245 20 137 30 144 60 77 90 50

The composition was determined by MLA analysis and XRF. The modal mineral content is shown in Table 4.2. Barite (24.6%), calcite (21.3%), dolomite (11.6%), and quartz (7.65%) were identified as the dominant gangue minerals, with bastnaesite (8.9%) and other rare earth minerals parisite (1.89%), monazite (0.99%), and allanite (0.28%), making up the valuable content of the ore. Once the software calculated the mineral content, it converted those numbers into elemental contents, which are compared with the semi-quantitative XRF values in Table 4.3.

The barium, cerium, and lanthanum values according to MLA are between two and four times as high as according to the XRF. The MLA sulfur concentration is slightly higher than that given by XRF. The calcium, iron, and silicon numbers agree generally well. The discrepancy between the methods, as well as the rise in REE content associated with grind time, is assumed to be due to an overestimation of the dense minerals, even though special care was taken to avoid such a bias in sample prep.

A false-color image of the largest size fraction of the unground ore is displayed in Figure 4.2, and one of the 90-minute ground sample 200 x 400 mesh size fraction is shown in Figure 4.3. The obvious particle size reduction, as well as the increase in liberation due to grinding, is evident upon comparison of the two images.

The bastnaesite does not strongly report to a specific size fraction, as can be seen from the mass distributions in Table 4.4. The bastnaesite grade tends to increase slightly with decreasing particle size and with grinding time. This is an unusual phenomenon, as it would be expected to remain constant because all samples were taken from the whole in a similar fashion – the only difference was the grinding time.

Table 4.2. Modal Mineral Content of Major Components of Bastnaesite Ore. REE-bearing minerals are in bold.

Mineral 762 µm 245 µm 144 µm 77 µm 50 µm Barite 24.6 26.4 26.1 27.3 30.9 Calcite 21.3 20.0 20.2 19.0 17.5 Dolomite 11.6 11.2 10.9 10.9 9.24 Bastnaesite 8.90 9.98 10.6 11.3 12.1 Quartz 7.65 7.95 7.66 7.38 6.64 K Feldspar 6.40 5.84 6.01 5.81 5.78 Parisite 1.89 2.06 2.20 2.29 2.46 Monazite 0.99 1.31 1.27 1.57 1.62 Allanite 0.28 0.24 0.21 0.21 0.15 SUM 83.6 85.0 85.2 85.8 86.4

Table 4.3. Elemental Composition of Bastnaesite Ore

MLA XRF Crushed Crushed Ore – 144 Ore – 144 Analyte 762 µm µm 50 µm 762 µm µm 50 µm Ba 14.5 15.4 18.2 5 7 8 Ca 11.7 11.1 9.6 10 13 13 Ce 3.71 4.41 5.04 2 2 2 Fe 2.75 2.7 2.41 2 2 3 La 3.61 4.33 4.97 0.8 1 1 Mg 1.82 1.65 1.39 3 3 3 S 3.64 3.83 4.51 2 2 2 Si 7.24 6.77 6.01 7 8 8 Sr 1.97 1.95 2.03 1 2 2 SUM 50.9 52.1 54.2 32.8 40.0 42.0

Figure 4.2. False-Color image of bastnaesite ore +50 mesh size fraction. Values represent surface area percentages.

Figure 4.3. False-Color image of 90-minute ground bastnaesite ore 200 x 400 mesh size fraction. Values represent surface area percentages.

The data in in Figure 4.4 and Table 4.5 shows that liberation increases with grind time for both mineral groups, with the greatest improvement coming after the first ten minutes of grinding. The crushed ore began with 22% of the REE minerals and 16% of the calcite fully liberated, reaching only 85% of the mass being more than 25% liberated for the REE minerals and 92% for the calcite. Grinding for 90 minutes improved the

52 liberation of the samples, upgrading to 71% and 69% of the minerals being 100% liberated, and almost all of the mass (98%) being more than 25% liberated.

Table 4.4. Mass and Bastnaesite Distributions. The mass percentage for each size fraction is given in plain text and the bastnaesite distribution in bold.

762 µm 245 µm 144 µm 77 µm 50 µm

+50 Mesh 60 56.9 ------50 X 100 14.4 13.5 34.7 30.5 8.1 4.2 0.2 0.1 0.1 0 Mesh

100 X 200 8.6 8.8 27.3 26.6 35.2 29.6 9.3 3.4 1.4 0.2 Mesh

200 X 400 5.7 6.4 13.1 12.7 20.3 19.5 33 31.4 23.1 15.8 Mesh

-400 Mesh 11.3 14.3 24.8 30.2 36.5 46.7 57.4 65.2 75.4 83.9

Total 100 100 100 100 100 100 100 100 100 100

Table 4.5. REE mineral and calcite liberation as cumulative mass recovery. Bolded values represent the combined total of REE minerals (bastnaesite, parisite, monazite, allanite). Percent Crushed Ore Liberated 762 µm 245 µm 144 µm 77 µm 50 µm 100 22 16 43 39 60 57 65 59 71 69 75 45 65 73 79 81 85 86 88 89 90 50 65 79 84 90 90 93 92 94 94 95 25 85 92 94 96 96 97 97 98 98 98

The zeta potential of the ore in distilled water was determined using a Stabino zeta potential device. The IEP of the ore was found to be 8.0 ± 0.3. IEP’s for barite, calcite, and bastnaesite were found to be 6.0 ± 0.2, 6.2 ± 0.4, and 6.0 ± 0.3. [50]

Figure 4.4. Cumulative REE mineral recovery by liberation class from MLA.

Figure 4.5. Molycorp ore zeta potential

The difference in IEP of the ore (Figure 4.5) from those of the minerals composing it suggests that while the main components are barite, calcite, and bastnaesite, the electrokinetic behavior of the ore is not similar to the behavior of any one major component.

4.2 Microflotation

Microflotation data was plotted in terms of elemental recovery and grade vs depressant concentration (or collector concentration when no depressant was used). Collector concentrations were held at either 5x10-5 M sodium oleate or 3x10-4 M octanohydroxamic acid for the depressant tests after a satisfactory response was obtained with those middle concentrations. Different oleate and hydroxamate concentrations were used after the literature survey showed that their effects were not similar at equal concentrations Oleate data is presented as filled points and hydroxamate data is shown with outlines. The experimental data for the flotation tests, including mass balances and elemental accountabilities, is given in APPENDIX C: EXPERIMENTAL DATA.

Reproducibility can be seen for the oleate system in comparing tests 2 and 13 and for the hydroxamate system in comparing tests H6 and H6b. These tests were performed using similar concentrations at carried out at different times to ensure that similar test conditions yielded repeatable results. However, the information from them should be considered qualitative at best.

Figure 4.6 and Figure 4.7 illustrate the difference between the collectors in terms of recovery and concentrate grade, respectively. Both show increases in recovery with concentration, but little difference in grade. The increase in recovery with concentration is expected: with more collector molecules in solution, more mass can be attached to bubbles to float.

Figure 4.6. Elemental recovery as a function of collector concentration

From Figure 4.7, it can be seen that the hydroxamate shows no more selectivity than the oleate with respect to the rare earth elements at the concentrations tested. A change in grade with respect to concentration would not be expected, but comparing the grades between the collectors shows no appreciable difference. The barium recovery increases in the hydroxamate system, but generally falls in the oleate system. No specific element’s grade was strongly affected.

Figure 4.7. Concentrate grade as a function of collector concentration.

Figure 4.8. Elemental recovery as a function of ammonium lignin sulfonate concentration.

The effect of ammonium lignin sulfonate is displayed in Figure 4.8 and Figure 4.9. Calcium and REEs show reduced recovery with increased concentration in both the oleate and hydroxamate systems, although the effect is stronger in the oleate system. The ammonium lignin sulfonate behavior suggests that it is acting in the same manner in both systems, although the effect is seen at a greater extent in the oleate system. This bolsters Gerdel and Smith’s proposition from [26] that it is acting on the surface of the mineral to inhibit collector adsorption.

Figure 4.9. Concentrate grade as a function of ammonium lignin sulfonate concentration.

The REE and calcium recoveries in both systems fell, but the hydroxamate system showed a stronger effect with respect to sodium silicate addition (Figure 4.10). Again, Figure 4.11 shows no appreciable change in grade.

The depression of hydroxamate flotation of all minerals by sodium silicate is attributed to the high pH reached during conditioning, as recovery is shown (Figure 2.22) to have a peak and sharp cliff at pH 10 and higher. The addition of sodium silicate raised the conditioning pH of these tests to 10.0, 10.6, and 10.8 (oleate); and 9.6, 10.1, and 10.2 (hydroxamate). Reduction of pH was not considered due to an effort of limiting reagent use and high acid consumption by reaction with calcium.

The cause for the continuous drop in hydroxamate recovery is likely due to a lack of collector adsorption, caused by the increase in pH by the sodium silicate. As the mechanism for adsorption relies on the reaction of collector molecules and charged hydroxy species, those species must be prevalent for flotation. But considering the dominance of uncharged Ce(OH)3 at pH of 10 and higher (shown previously in Figure 2.3), the adsorption process breaks down. The result is reduced recovery. Reduction of pH prior to conditioning was not used in an effort to keep experimental conditions simple for these qualitative tests.

Figure 4.10. Elemental recovery as a function of sodium silicate concentration.

Figure 4.11. Concentrate grade as a function of sodium silicate concentration.

Whereas sodium silicate affected the hydroxamate system more strongly, copper nitrate affected the recoveries in the oleate system (Figure 4.12). All species’ recoveries fell. The effect on rare earth grade was minimal (Figure 4.13).

Figure 4.12. Elemental recovery as a function of copper (II) nitrate concentration.

Figure 4.13. Concentrate grade as a function of copper (II) nitrate concentration.

Copper nitrate addition to the system appears to precipitate the fatty acid collector, based on the decreased recovery in the oleate system and not the hydroxamate system. This is in line with Gaudin’s decreased recovery findings from Figure 2.24, and is merely the result of collector precipitation by the metal salt. Hydroxamate flotation was unaffected.

The results show that gaining separation in flotation is difficult. The adsorption of ammonium lignin sulfonate onto mineral surfaces non-selectively decreased recovery. The addition of sodium silicate did not have an effect on oleate flotation, and increased the pH too much for the charged hydroxy species to interact with the hydroxamate collector. Further buffering the reagent could prevent this. The copper sulfate ions did not interact with hydroxamate flotation, but precipitated the oleate collector. Thus, it can be shown that selectivity is difficult to achieve by the chemical processes in flotation alone.

4.3 Magnetic Separation

TREE grade and recovery data were plotted for the six Wet High Intensity Magnetic Separation (WHIMS) tests that were conducted. In Figure 4.14 it can be seen that the TREE grade increased with finer grind size and higher field strength. The figure also shows that TREE recovery is greatest with the coarsest feed and the highest field strength. Such a result is expected: as the bastnaesite is magnetized in the separator, anything that is associated with it will report to the magnetic fraction. When very little is attached (because the grind size is fine), the magnetic product TREE grade is high – but when the grind size is coarse, the associated particles also report to that fraction – and so the coarse grade is low. A similar phenomenon would cause the high recovery in the coarse product – fine bastnaesite grains were able to pass through the separator attached to relatively large nonmagnetic grains.

The distribution of the barium between the nonmagnetic and the magnetic fractions is noteworthy. Figure 4.15 displays that the barium content in the nonmagnetic fraction is much higher than that in the magnetic fraction. Although the recoveries to the magnetic product only ranged from 3-17% (see Table 4.6), the potential exists for further optimization of this process to result in substantial rejection of barite from the feed.

Table 4.6. Calcium, barium, and TREE grade and recovery of WHIMS concentrate products.

Field Strength, Grade, Percent Recovery, Percent

Run P80 , µm Gauss Ca Ba TREE Ca Ba TREE Mass 1 762 5000 10 4 2.4 7.1 5.4 10.3 8.8 2 50 5000 8 2 4.9 1.5 0.6 2.9 2.8 3 762 10000 11 3 3.7 17.0 9.3 28.5 16.8 4 144 7500 11 2 3.8 2.6 0.8 4.1 4.0 5 50 10000 14 2 8.4 5.9 1.6 11.5 6.6 6 144 7500 12 3 9.6 3.3 1.3 7.4 4.6

Figure 4.14. WHIMS concentrate TREE grade (left)and recovery (right).

Figure 4.15. Barium distribution between WHIMS magnetic and non-magnetic fractions.

4.4 Gravity Concentration

The float and sink product calcium grades and mass recoveries for the heavy liquid tests are given in Figure 4.16 and Figure 4.17. The most striking result is that the rare earths and barium almost exclusively report to the sink product, no matter the grind time or specific gravity of the fluid. The calcium trends are shown in Figure 4.16. Calcium recovery to the sinks increases with the density of the fluid and with increasing grinding time (i.e. finer size). The sink product recovery decreases with grinding time and increasing specific gravity. These trends do not extend to the highest density fluid, likely because the high viscosity of the fluid interfered with the settling of the particles. The complete mass and elemental balance for each test can be found in EXPERIMENTAL DATA.

Figure 4.16. Calcium recoveries to HLA products.

These results were as expected. The concentration criteria (CC from Table 2.4) for bastnaesite and barite are 2.35 and 2.06. These represent relatively good separation. Unsurprisingly, these are the minerals that reported to the sink product. Calcite has a criterion of 1.00, which dictates that a separation in water is fundamentally impossible. As the fluid became denser, the density difference pushed more of the calcite to the floats.

For a non-selective process, the mass of each mineral would match that of the overall mass recovered. If the mineral reports to a product at a greater percentage than the product represents from the feed, it can be deduced that a preferential separation exists. Such is the case with barium and the rare earths. The increased liberation with grind time and low density of calcite, the main calcium-containing mineral in this ore, explain the increased recovery to the floats at high fluid densities.

This trend continued in the concentrate streams on the Wilfley shaking table. The reprocessed ore showed good separation into different streams (Figure 4.18). The compositions of spots 1, 2, 3, and 4 are given in

Table 4.7 below. These concentrations show the motion of the table is successful in separating the heavier bastnaesite and barite from the lighter calcite. Unfortunately, damage to the table on the discharge end destroyed the fluid film and prevented effective separation of a concentrate from a middling product. As the minerals flowed

63 across the table and began separating according to their density into different streams, the streams remixed ahead of the dry spot, eliminating that separation.

Figure 4.17. Mass recovery as a function of fluid density.

Table 4.7. Stream Concentrations

Spot # (1) (2) (3) (4) Ca 1.6 1.0 1.4 29 Ba 15 35 40 1.0 TREE 39 27 20 1.2

The Deister table was used to generate several mass balances. Table 4.8 compares two similar table trials. Table 4.9 shows the results of washing the concentrate twice after the first separation. Table 4.8 shows consistent reduction of the calcium grade in the concentrate, and an increase in the REE grade. The recoveries of the rare earths in the concentrates are still low, at 23% and 31%; but they are close to double the calcium recoveries of 8% and 8%. In Table 4.9 it can be seen that 97% of

64 the calcium was rejected from the reprocessed concentrate (as only 3% was recovered).

Figure 4.18. Shaking Table Streams

Table 4.8. Shaking table mass balance for two similar 500 g tests, 2 and 3.

Feed (Calculated) Concentrate MassGrade (%) Mass Grade (%) Recovery (%) Test (g) Ca Ba TREE (g) Ca Ba TREE Mass Ca Ba TREE 2 500 13 14 4.4 53 12 16.18 7.8 11 9 12 19 3 500 13 16 5.1 63 9.3 20 9.2 13 9 16 23 Tailings Mass Grade (%) Recovery (%) (g) Ca Ba TREE Mass Ca Ba TREE 2 All recoveries correspond 180 13.9 13.8 3.4 36 37 35 28 3 to a 500 g feed 160 13.8 14 3.4 32 35 29 22

This suggests the possibility of a roughing – cleaning – scavenging circuit to enrich the concentrate while retrieving as much rare earth content as possible. That proposed circuit was tested by re-tabling the initial concentrate, middlings, and tails for a total of eight products (no mass reported to the tails product after reprocessing the initial

65 concentrate). The weight percentages and total recoveries with respect to the initial feed of each of the eight products of the multi-table circuit can be found in APPENDIX C: EXPERIMENTAL DATA. That table also includes extended versions of Table 4.8 and Table 4.9, including the middling product and mass balances.

From the 9.2 wt% TREE concentrate (of Test 3), a rare-earth-enriched concentrate (20%) was created (Test 6). Of the rare earths in the initial feed, 25% reported to the second concentrate. As only 2% of the calcium gangue reported to the cleaned concentrate, that product has the best rare earth recovery and lowest calcium content.

The middling from Test 3 held 38% of the rare earths and produced (through Test 5) a concentrate, middling, and tails with elevated calcium contents (17%, 16%, and 15%) and low rare earth contents (2.5%, 3.4%, and 3.5%). The significant mass of the middling products means that those products still contained 3%, 4%, and 21% of the rare earths.

Slimes are likely responsible for carrying 5% of the rare earth content into the tails of the Test 3 tails (during test 4). Only 7% of the REEs reported to that concentrate, and 19% ended up in its middlings. The concentrate had slightly enriched rare earth content, at 5.4%, but the middlings and tails only had grades of 1.9% and 3.2% and high calcium contents (13%, 17%, and 15%).

Based on the grades and recoveries of the different products, a flowsheet (Figure 4.19) might contain roughing tables to reject slimes, cleaning tables to produce a final concentrate, and scavenging tables to upgrade the middling product. Slimes production could be controlled in the comminution circuit by raising the grind size, but that would reduce the liberation factor of the ore. A rod mill (which limits creation of fines in grinding) and cyclones (for particle classification, to reduce overgrinding of fine particles) are already employed (Figure 1.1).

Table 4.9. Mass balance for thrice-tabled concentrate. Products are the final concentrate and combined total middlings and tailings.

Feed (Calculated) Concentrate MassGrade (%) Mass Grade (%) Recovery (%) (g) Ca Ba TREE (g) Ca Ba TREE Mass Ca Ba TREE 500 13 15 5.3 26 8 21 14 5 3 7 13

Middlings Recoveries correspond to 500 g feed Mass Grade (%) Recovery (%) (g) Ca Ba TREE Mass Ca Ba TREE 314 15.5 9 3 63 67 51 50

Tailings Balance (%) Mass Grade (%) Recovery (%) Mass Ca Ba TREE (g) Ca Ba TREE Mass Ca Ba TREE 93.32 93.32 93.32 93.32 127 13.5 14 3.6 25 27 23 17

Figure 4.19. Gravity concentration circuit.

Six responses of the Falcon test campaign were recorded: TREE grade, calcium grade, barium grade, TREE recovery, calcium recovery, and barium recovery. These responses were analyzed with Design Expert 9. The factorial design of the experiments allowed for the analysis of the significant effects on the process. Only two responses produced a signal-to-noise ratio high enough (that is, greater than 4.0) to consider navigation of the design space: TREE grade and calcium recovery (Table 4.10).

However, both models (TREE grade and calcium recovery) showed that all of the input variables were statistically insignificant, and that the models themselves were insignificant. Their R2 values were .7757 and .7527, respectively. The Analysis of Variance (ANOVA) table for the TREE grade model is given in Table 4.11.

Table 4.10. Signal-to-Noise Ratio for Falcon DOE Responses

Response Signal-to-Noise Ratio TREE Recovery 1.61 TREE Grade 6.65 Ca Recovery 4.15 Ca Grade 2.43 Ba Recovery 2.32 Ba Grade 2.14

Table 4. 11 . ANOVA Table for the TREE Grade Model.

Sum of Mean F p-value Source Squares df Square Value Prob > F Model 6.45 5 1.29 3.46 0.0998 A-G-Force 0.18 1 0.18 0.48 0.5181 B-Feed Rate 1.13 1 1.13 3.02 0.1429 C-Feed Size 2 1 2 5.36 0.0684 AB 0.72 1 0.72 1.93 0.2233 BC 2.42 1 2.42 6.49 0.0514 Residual 1.86 5 0.37 Lack of Fit 1.42 3 0.47 2.12 0.3369 Pure Error 0.45 2 0.22 2 Cor Total 8.31 10 R = 0.7757

The grade and recovery from each test is given in Table 4.12. Grade is most strongly a function of feed size and feed rate, with G-Force having a smaller effect. The best results came from the low G-Force, low feed rate, and small feed size tests. Tests

68 were repeated on those optimum conditions, as they were the ones predicted by the design model to give the best grade, and those results are given in Table 4.12.

Table 4.12. Grades, recoveries, and concentration ratios of Falcon products.

DOE Falcon G-Force Grind Time Feed Rate Recovery TREE Grade Concentration Standard Run G's minutes kg/hr Percent Percent Ratio

Opt 1 100 90 30 50.6 8.7 2.56 Opt 2 100 90 30 48.1 8.0 2.35 Opt 3 100 90 30 43.9 6.8 2.00 1 1 100 90 30 36.8 8.5 2.50 2 8 250 90 30 38.8 7.4 2.18 4 3 250 90 60 52.3 6.3 1.85 8 6 250 30 60 40.7 6.3 1.85 5 9 100 30 30 44.4 6.2 1.82 9 5 175 60 45 36.5 6.1 1.79 7 10 100 30 60 37.6 6.1 1.79 3 4 100 90 60 53.6 5.9 1.74 11 7 175 60 45 59.2 5.9 1.74 6 2 250 30 30 40.9 5.5 1.62 10 11 175 60 45 30.9 5.2 1.53

1.000

0.800

0.600

0.400

0.200 D esirability

0.000

30 37.5 45 100 137.5 B: Feed Rate (kg/hr) 52.5 175 212.5 60 250 A: G-Force (G's)

Figure 4.20. Desirability Surface for Falcon tests with 50-micron feed size.

All three tests produced very high grades, affirming the optimum conditions, predicted by the desirability surface in Figure 4.20. The rare earth and barium recoveries double that of the mass recovery, pointing toward the same preferential separation as the heavy liquid work. The combined grade of these products is also approximately double that of the feed. Concentration ratios for all tests are given in Table 4.12.

4.5 Bench Flotation

As shown in Table 4.13, rare earth grade remained the same, but recovery improved for the gravity preconcentrate flotation test as compared to the ground ore test. Calcium recovery was similar, but the grade was reduced.

Table 4.13. Bench Flotation Results: TREE and Calcium Grade and Recovery

TREE Ca Grade TREE Recovery Ca Grade Recovery Test (Percent) (Percent) (Percent) (Percent)

P80 = 45 μm 30 77 9 10 Gravity Con 30 82 5 10

Although the gravity preconcentrate yielded a similar grade and higher recovery for the flotation unit operation, it must be noted that the current gravity recovery is only 50%, meaning overall recovery is only 41%. To eclipse the reported flotation performance from Molycorp, the process must generate a combined gravity and flotation recovery of greater than 65%. At this point, it does not. However, looking at the stepwise recovery and grade from each pass through the Falcon (given in Figure 4.21), it can be estimated that a sufficient recovery may be produced at an improved grade with three more passes through the Falcon. This estimation is a linear forecast of the average grades and recoveries of the first three steps.

Figure 4.21. TREE Grade vs Recovery of Falcon Concentrates.

CHAPTER 5: ECONOMIC ANALYSIS

Two separate economic models were developed: one utilizing a froth flotation circuit for concentration and one using gravity preconcentration followed by a froth flotation circuit. The features of each are given in Table 5.2. Both conceived plant models treat 400,000 tonnes per year of ore, 8% of which is rare earth oxides. They have the same comminution circuit in front of the concentrator containing a jaw crusher, SAG mill, ball mill, and regrind mill. The flotation circuits use 18 rougher, 18 cleaner, and 36 scavenger cells. Both models’ concentrate products are assumed to be upgraded by cleaner flotation to 60% REO while keeping the same calcium contents from the rougher flotation. A hydrochloric acid leach is the next step, where it is assumed that the acid completely reacts with the calcium in the ore (from calcite), according to the reaction

Equipment sizes and costs were based off those given in the 1,000 tonne/day Flotation Mill model in CostMine 2014 – the closest model in magnitude to a 400,000 tonne/year operation. Reagent, labor, and energy costs were also sourced from CostMine or estimated from vendor information. Capital expenses were derived from the formulas given by Mular. [51] Recent rare earth oxide prices were obtained from the Metal Pages website. The revenues, profits, and expenses in the section are pre-tax values.

The first model uses a circuit of flotation cells to recover 65% of the REO content of the ore (which made up 8% of the total ore body) using a fatty acid collector. The mine produces 400,000 tons/year of ore. The recovered 20,800 tonnes of rare earth oxides generate annual profit of $69,390,000 using a $4,000/tonne REO price. The second model places a series of shaking tables and centrifugal concentrators in front of the froth flotation circuit, this time utilizing a hydroxamic acid collector. The same amount and grade of ore is mined but, reflecting the improvement shown by the bench flotation tests, this design simulates 82% REO recovery. The increased production revenue and lowered reagent cost (mainly from the acid) yields an annual profit of

$91,754,000. The components of this $22,364,000 difference are displayed in Table 5.1. The annual labor and energy expenses are higher due to the implementation of the gravity circuit, but the decrease in reagent expense (because less acid is consumed) lead to an annual operating expense that would offset the increased (one-time) capital expense in about five years. These models predict a 10-year pre-tax NPV at an 8% discount rate of $311,691,000 and $413,919,000 for the flotation and gravity-flotation plants, respectively.

These plants have payback periods of approximately four months and produce internal rates of return of 358% and 412%, respectively. Detailed charts of the revenues, expenses, cash flows, net present values, and internal rates of return are given in ECONOMIC ESTIMATES.

Table 5.1. Comparison of Plant Annual and Capital Expense Differences (calculated as gravity-flotation plant values less flotation plant values).

Rare Earth Flotation Concentrator, Inc. Rare Earth Gravity-Flotation Concentrator, Inc. Difference Profit$ 69,390,272 $/year Profit $ 91,754,116 $/year Profit $ 22,363,843 $/year Production $ 83,200,000 $/year Production $ 104,960,000 $/year Production $ 21,760,000 $/year Ore Mined 400,000 tonnes/year Ore Mined 400,000 tonnes/year Ore Mined 400,000 tonnes/year REO Grade 8% Percent REO Grade 8% Percent REO Grade 8% Percent REO Recovery 65% Percent REO Recovery 82% Percent REO Recovery 17% Percent REO Price 4,000 $/tonne REO Price 4,000 $/tonne REO Price 4,000 $/tonne Con REO Grade 60% Percent Con REO Grade 60% Percent Con REO Grade 60% Percent Con Ca Grade 9% Percent Con Ca Grade 5% Percent Con Ca Grade 4% Percent Operating Expenses $ 13,809,728 $/year Operating Expenses $ 13,205,884 $/year Operating Expenses $ (603,843) $/year Labor $ 3,485,720 $/year Labor $ 3,770,420 $/year Labor $ 284,700 $/year Reagents $ 9,403,325 $/year Reagents $ 8,496,108 $/year Reagents $ (907,217) $/year Fatty Acid (Collector) $ 3.00 $/kg Hydroxamic Acid (Collector) $ 6.00 $/kg Collector $ 3.00 $/kg 0.3 kg/tonne Ore 0.3 kg/tonne Ore 0.3 kg/tonne Ore Ammonium Lignin Sulfonate $ 3.00 $/kg Ammonium Lignin Sulfonate $ 3.00 $/kg Ammonium Lignin Sulfonate $ 3.00 $/kg 2.5 kg/tonne Ore 2.5 kg/tonne Ore 2.5 kg/tonne Ore Hydrochloric Acid $ 250 $/tonne Hydrochloric Acid $ 250 $/tonne Hydrochloric Acid $ 250 $/tonne 42.36 kg/tonne Ore 29.69 kg/tonne Ore 6.16 kg/tonne Ore Energy $ 920,683 $/year Energy $ 939,357 $/year Energy $ 18,673 $/year Capital Expenses $ 19,318,444 Capital Expenses $ 22,254,940 Capital Expenses $ 2,936,495

Sensitivity analysis of the gravity-flotation plant shows that production output and price have much more significant effects on the NPV than the operating and capital expenses. This is the primary reason for the economic favorability of the second model. The enormous increase in revenue from production quickly makes up for the slight decrease due to higher capital and operating expenses.

The second model is built assuming an idealized 82% overall REO recovery. The actual recovery based on this work would only be 41%, which would not be as attractive as the fatty acid flotation achieving 65% recovery. Enhancing the Falcon recovery to 78% and maintaining the flotation recovery of 82% would produce an overall

72 recovery of 64%. This recovery value is the break-even point at which the gravity- flotation plant would generate a higher ten-year NPV than the fatty acid flotation plant.

Figure 5.1. Sensitivity Analysis for the gravity-flotation concentrator.

Table 5.2. Concentrator economic model parameters (Operation).

Rare Earth Flotation Concentrator, Inc. Rare Earth Gravity-Flotation Concentrator, Inc.

Production $ 83,200,000 $/year Production $ 104,960,000 $/year Ore Mined 400,000 tonnes/year Ore Mined 400,000 tonnes/year REO Grade 8% Percent REO Grade 8% Percent REO Recovery 65% Percent REO Recovery 82% Percent REO Price 4,000 $/tonne REO Price 4,000 $/tonne Con REO Grade 60% Percent Con REO Grade 60% Percent Con Ca Grade 9% Percent Con Ca Grade 5% Percent Working Hours 24 hours/day Working Hours 24 hours/day Working Days 365 days/year Working Days 365 days/year Schedule 8,760 hours/year Schedule 8,760 hours/year Operating Expenses $ 13,809,728 $/year Operating Expenses $ 13,205,884 $/year Labor $ 3,485,720 $/year Labor $ 3,770,420 $/year Mill Superintendent $ 131,100 $/year Mill Superintendent $ 131,100 $/year Maintenance Foreman $ 124,200 $/year Maintenance Foreman $ 124,200 $/year Plant Foreman (3) $ 110,400 $/year Plant Foreman (3) $ 110,400 $/year Metallurgist $ 113,200 $/year Metallurgist $ 113,200 $/year Process Technician $ 82,800 $/year Process Technician $ 82,800 $/year Instrument Technician $ 85,600 $/year Instrument Technician $ 85,600 $/year Process Foreman $ 103,500 $/year Process Foreman $ 103,500 $/year Crusher Operator $ 34.43 $/hour Crusher Operator $ 34.43 $/hour 2 workers/day 2 workers/day Grinding Operator $ 32.50 $/hour Grinding Operator $ 32.50 $/hour 3 workers/day 3 workers/day Gravity Operator $ 32.50 $/hour 3 workers/day Flotation Operator $ 32.50 $/hour Flotation Operator $ 32.50 $/hour 3 workers/day 3 workers/day Leaching Operator $ 32.50 $/hour Leaching Operator $ 32.50 $/hour 3 workers/day 3 workers/day Assayer $ 34.50 $/hour Assayer $ 34.50 $/hour 1 worker/day 1 worker/day Samplers $ 24.98 $/hour Samplers $ 24.98 $/hour 3 workers/day 3 workers/day Laborers $ 23.25 $/hour Laborers $ 23.25 $/hour 6 workers/day 6 workers/day Mechanics $ 36.11 $/hour Mechanics $ 36.11 $/hour 3 workers/day 3 workers/day Electricians $ 36.23 $/hour Electricians $ 36.23 $/hour 3 workers/day 3 workers/day Reagents $ 9,403,325 $/year Reagents $ 8,496,108 $/year Soda Ash $ 1.75 $/kg Soda Ash $ 1.75 $/kg 2.5 kg/tonne Ore 2.5 kg/tonne Ore MIBC (Frother) $ 2.86 $/kg MIBC (Frother) $ 2.86 $/kg 0.05 kg/tonne Ore 0.05 kg/tonne Ore Fatty Acid (Collector) $ 3.00 $/kg Hydroxamic Acid (Collector) $ 6.00 $/kg 0.3 kg/tonne Ore 0.3 kg/tonne Ore Ammonium Lignin Sulfonate $ 3.00 $/kg Ammonium Lignin Sulfonate $ 3.00 $/kg 2.5 kg/tonne Ore 2.5 kg/tonne Ore Hydrochloric Acid $ 250 $/tonne Hydrochloric Acid $ 250 $/tonne 42.36 kg/tonne Ore 29.69 kg/tonne Ore Energy $ 920,683 $/year Energy $ 939,357 $/year Electrical Power 0.089 $/kWh Electrical Power 0.089 $/kWh Comminution 21.448 kWh/tonne Comminution 21.448 kWh/tonne Crusher 150.00 HP/unit Crusher 150.00 HP/unit SAG Mill 500.00 HP/unit SAG Mill 500.00 HP/unit Ball Mill 1,000.00 HP/unit Ball Mill 1,000.00 HP/unit Regrind Mill 150.00 HP/unit Regrind Mill 150.00 HP/unit Grinding Slurry Pumps 10.00 HP/unit Grinding Slurry Pumps 10.00 HP/unit Gravity Concentration 0.52 kWh/tonne Tables 3.00 HP/unit Centrifugal Concentrator 7.50 HP/unit Flotation 4.298 kWh/tonne Flotation 4.298 kWh/tonne Rougher Flotation Cells 5.00 HP/unit Rougher Flotation Cells 5.00 HP/unit Cleaner Flotation Cells 5.00 HP/unit Cleaner Flotation Cells 5.00 HP/unit Scavenger Flotation Cells 3.00 HP/unit Scavenger Flotation Cells 3.00 HP/unit Flotation Slurry Pumps 2.50 HP/unit Flotation Slurry Pumps 2.50 HP/unit Concentrate Slurry Pumps 1.00 HP/unit Concentrate Slurry Pumps 1.00 HP/unit Tailings Slurry Pump 25.00 HP/unit Tailings Slurry Pump 25.00 HP/unit Leaching 0.12 kWh/tonne Leaching 0.12 kWh/tonne Leaching Slurry Pump 1.00 HP/unit Leaching Slurry Pump 1.00 HP/unit

Table 5.3. Concentrator economic model parameters (Capital Costs).

Capital Expenses $ 19,318,444 Capital Expenses $ 22,254,940 76 cm x 107 cm Double Toggle Ja\w Crusher $ 281,100 $/unit 76 cm x 107 cm Double Toggle Ja\w Crusher $ 281,100 $/unit 1 unit 1 unit 4.9 m x 1.5 m SAG Mill $ 1,375,000 $/unit 4.9 m x 1.5 m SAG Mill $ 1,375,000 $/unit 1 units 1 units 3.0 m x 5.5 m Ball Mill $ 1,243,600 $/unit 3.0 m x 5.5 m Ball Mill $ 1,243,600 $/unit 1 units 1 units 2.1 m x 2.3 m Regrind Ball Mill $ 514,800 $/unit 2.1 m x 2.3 m Regrind Ball Mill $ 514,800 $/unit 1 units 1 units 757 lpm Grinding Slurry Pump $ 10,200 $/unit 757 lpm Grinding Slurry Pump $ 10,200 $/unit 4 units 4 units 46 cm Cyclone $ 9,290 $/unit 46 cm Cyclone $ 9,290 $/unit 3 units 3 units 7.53 mpth (coarse sands) Triple Deck Shaking Table $ 64,750 $/unit 10 units 72.6 mtph Centrifugal Gold Concentrator $ 110,900 $/unit 2 units 1.4 cu m Rougher Flotation Cell $ 27,500 $/unit 1.4 cu m Rougher Flotation Cell $ 27,500 $/unit 18 units 18 units 1.4 cu m Cleaner Flotation Cell $ 27,500 $/unit 1.4 cu m Cleaner Flotation Cell $ 27,500 $/unit 18 units 18 units 0.71 cu m Scavenger Flotation Cell $ 27,500 $/unit 0.71 cu m Scavenger Flotation Cell $ 27,500 $/unit 36 units 36 units 189 lpm Flotation Slurry Pump $ 8,745 $/unit 189 lpm Flotation Slurry Pump $ 8,745 $/unit 14 units 14 units 76 lpm Concentrate Slurry Pump $ 8,745 $/unit 76 lpm Concentrate Slurry Pump $ 8,745 $/unit 3 units 3 units 1,893 lpm Tailings Slurry Pump $ 13,330 $/unit 1,893 lpm Tailings Slurry Pump $ 13,330 $/unit 1 units 1 units 3.66 m x 3.66 m (h x d) Leaching Tanks $ 22,500 $/unit 3.66 m x 3.66 m (h x d) Leaching Tanks $ 22,500 $/unit 3 units 3 units 76 lpm Leach Slurry Pump 8,745 $/unit 76 lpm Leach Slurry Pump 8,745 $/unit 3 units 3 units Capital Equipment (CE)$ 5,718,900 Capital Equipment (CE)$ 6,588,200 Delivery (0.03 x CE) $ 171,567 Delivery (0.03 x CE) $ 197,646 Installation (0.20 x CE) $ 1,143,780 Installation (0.20 x CE) $ 1,317,640 Piping (0.16 x CE) $ 915,024 Piping (0.16 x CE) $ 1,054,112 Electrical (0.19 x CE) $ 1,086,591 Electrical (0.19 x CE) $ 1,251,758 Instrumentation (0.075 x CE) $ 428,918 Instrumentation (0.075 x CE) $ 494,115 Buildings (0.525 x CE) $ 3,002,423 Buildings (0.525 x CE) $ 3,458,805 Plant Services (0.11 x CE) $ 629,079 Plant Services (0.11 x CE) $ 724,702 Site Improvements (0.105 x CE) $ 600,485 Site Improvements (0.105 x CE) $ 691,761 Field Expenses (0.11 x CE) $ 629,079 Field Expenses (0.11 x CE) $ 724,702 Project Engineering and Construction (0.31 x $ 1,772,859 $ 2,042,342 CE) Project Engineering and Construction (0.31 x CE) Fixed Capital (FC)$ 16,098,704 Fixed Capital (FC)$ 18,545,783 Working Capital (20% of FC)$ 3,219,741 Working Capital (20% of FC)$ 3,709,157 Sustaining Capital$ 166,667 $/year Sustaining Capital$ 166,667 $/year

CHAPTER 6: CONCLUSIONS

Cursory microflotation studies showed that although preferential selectivity using hydroxamate as compared to oleate collectors was suggested by single-mineral microflotation studies, the same result was not achieved when using an ore mainly comprised of those minerals. Typical flotation depressants were unable to improve the selectivity in the system under the conditions explored. While ammonium lignin sulfonate depressed calcite, it also depressed bastnaesite, leading to no appreciable change in grade. Sodium silicate raised the pH too high for hydroxamate adsorption. Copper (II) nitrate addition to the system had no effect on the hydroxamate flotation.

Heavy liquid analysis and gravity concentration trials have shown that the difference in specific gravities of the minerals in this ore can be exploited to effect a separation. The calcium-containing minerals (predominately calcite) floated at higher fluid densities, while the rare earth minerals always sank. This effect was also seen in the Falcon Concentrator, with the best results (51% TREE recovery) showing that the longest-ground, most-liberated samples had the highest concentration of rare earths (8.7% TREE). Qualitative demonstrations on the shaking table suggest it could be used once the proper conditions have been determined.

Magnetic separation was briefly explored. Barite rejection from the concentrate was seen in the few tests performed, with an average barium recovery to the magnetic fraction of 3.1%. However, the mass recoveries of the magnetic fraction averaged around 7% and showed little change in TREE grade, so no further testing was performed.

Comparative flotation showed that a 30% TREE, 9% calcium concentrate could be produced at 77% TREE recovery using hydroxamic acid. Addition of a gravity preconcentration step reduced the calcium content to 5% while obtaining TREE grades and recoveries of 30% and 82%, respectively. While flotation of the gravity feed outperformed flotation of the ground ore feed, the combined process underperformed. Increasing the number of passes through the Falcon from three to six was estimated to generate a sufficient concentrate.

Economic analysis showed that placing a gravity preconcentration circuit in front of froth flotation will quickly recover the capital expense and yield a higher annual profit. The internal rates of return for both estimates were qutie high. The increase in recovery and reduction in downstream costs of leaching were the drivers of this difference. The factors to which the model is most sensitive were the level of production and the price of the rare earth oxides.

CHAPTER 7: SUGGESTIONS FOR FUTURE WORK

Future work should focus on size-classified centrifugal concentration, to see if splitting the feed into a coarse and fine fraction would increase overall recovery. The optimum slurry density was not determined in this work. Plants especially conscious of their water balance may wish to investigate that factor with a more robust experimental setup that is resistant to sanding.

Magnetic separation showed evidence of barite rejection, but at uninspiring mass recoveries. If the mass recovery was improved, perhaps with more passes through the WHIMS, and it was shown that barite is truly rejected in substantial quantities, flotation depressant requirements could be reduced.

A combination of gravity, magnetic, and flotation unit operations should also be investigated to improve plant throughput and reduce flotation reagent consumption. Gravitational preconcentration operations have shown the ability to upgrade the rare earth content of the flotation feed, but must be optimized to do so without sacrificing overall recovery. Together these operations would reduce throughput in the flotation circuit, lower reagent costs, and improve overall plant grade and recovery.

Once the concentration of the ore by the above methods has been completed, more surface chemistry analysis should be performed. The zeta potential and adsorption density of the ore should be investigated before and after concentration, to determine the effect of reducing the size and removing the calcite. The interaction of the collectors and depressants on the fine-sized, calcite-free sample could inform on future reagent selections.

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APPENDIX A: FIRST LOT MINERALOGY

Figure A.1. Cumulative particle size analysis of ground Molycorp bastnaesite ore.

Table A.1. Size fraction volume and weight percentages

Microtrac Sieve Analysis Size Fraction vol % wt % +100 3.11 0.63 100x325 49.36 43.61 325x400 4.97 3.77 -400 42.56 51.99

Table A.2. Composition (weight percent) of each size fraction as reported by MLA analysis.

Mineral Formula 100 100 x 325 325 x 400 Barite BaSO4 13 29.9 41.5 Calcite CaCO3 11.2 17.8 14.2 Bastnaesite (Ce,La)(CO3)F 10.6 10.8 12.8 Quartz SiO2 17.4 6.78 4.02 K_Feldspar KAlSi3O8 7.78 2.63 1.63 FeO Fe3O4 2.54 2.59 2.06 Parisite Ca(Ce,La)2(CO3)3F2 1.87 2.26 2.74 Ankerite CaFe(CO3)2 2.08 1.91 2.16 Chlorite (Mg3,Fe2)Al(AlSi3)O10(OH)8 2.31 1.62 1.23 Strontianite SrCO3 1.02 1.31 1.31 Monazite (La,Ce)PO4 0.69 1.2 1.84 Biotite K(Mg,Fe)3(AlSi3O10)(OH)2 3.17 1.17 0.67 Apatite Ca5(PO4)3F 0.12 0.12 0.12 Kaolinite Al2Si2O5(OH)4 0.11 0.29 0.17 Ilmenite FeTiO3 0.07 0.02 0.04 Rutile TiO2 0.1 0.11 0.28 Mica KAl2(AlSi3O10)(OH)2 6.62 1.88 1.13

Plagioclase (Na,Ca)(Al,Si)4O8 3.17 0.65 0.38 Allanite (Ca,Ce)2(Al,Fe)3(SiO4)(Si2O7)O(OH) 0.14 0.23 0.37 Hornblende (Ca2,Na)(Mg2FeAl)Si6O22(OH)2 0.25 0.16 0.2 Pyrite FeS2 0.13 0.11 0.04 Zircon ZrSiO4 P 0.08 P Magnetoplumbite Pb(Fe,Mn)12O19 0.01 0.02 0.06 Galena PbS 0.01 0.02 0.06 Enargite_Tenn Cu3AsS4 ND 0.01 P Thorite ThSiO4 P 0.01 0 Celestine SrSO4 0.46 0.81 1.1 Dolomite CaMg(CO3)2 15.2 15.6 9.92

Figure A.2. Liberation of Bastnaesite according to size fraction. Left: +100 mesh, Center: 100 x 325 mesh, Right: 325 x 400 mesh.

Figure A.3. False-color image showing particle association of the +100 mesh size fraction.

Figure A.4. False-color image showing particle association of the 100 x 325 mesh size fraction.

Figure A.5. False-color image showing particle association of the 325 x 400 mesh size fraction.

Figure A.6. Measured and WPPF-calculated diffractograms and residual plots for the bastnaesite ore.

Figure A.7. Bastnaesite ore diffractogram with candidate phases.

Table A.3. Composition (mass percent) of each size fraction as reported by QEMSCAN.

Mineral +100 100x325 325x400 -400 Barite 12.9 35.75 43.42 44.7 Calcite 10.3 18.8 16.23 16.6 Bastnaesite-(Ce) 8.14 9.31 10.29 5.21 Quartz 19 6.74 4.13 3.42 K-Feldspar 16.9 4.15 2.5 2.24 Fe Oxides 2.05 1.59 1.28 1.67 Parisite-Synchysite 2.58 1.79 2.02 1.35 Ankerite/(FeOx +Calcite) 7.8 7.56 5.3 4.14 Chlorite 3.37 2.47 1.94 1.95 Strontianite 0.78 0.76 0.75 0.31 Monazite-(Ce) 0.88 1.31 1.89 2.27 Biotite 5.78 2.9 1.64 1.99 Apatite (F) 0.35 0.29 0.6 0.63 Kaolinite 0.26 0.37 0.3 0.46 Ilmenite 0.09 0.09 0.08 0.1 Rutile/Anatase 2.42 3.09 4.93 9.61 Sericite 5.33 1.6 1.27 1.51

Mn Oxides 0.18 0.65 0.49 0.58 Cerianite + Quartz 0.5 0.27 0.3 0.46 Sulfur trap 0.01 0.17 0.08 0.07 Ancylite-possible Si-interference 0.19 0.11 0.15 0.2 Xenotime 0.04 0.08 0.18 0.24 Ce-rich Mn + Fe oxides (>6wt.% Ce) 0.06 0.06 0.05 0.03 Ancylite 0.06 0.04 0.07 0.14 Alunite 0.03 0.03 0.05 0.1 Fe-sulfur trap 0.08 0.03 0.05 0.08

Table A.4. Overall calculated mineral abundance of the ground Molycorp bastnaesite ore.

Weight Mineral Percent Barite 40.54 Calcite 17.48 Bastnaesite 7.21 Quartz 4.99 K_Feldspar 3.17 FeO 1.62 Parisite 1.57 Ankerite 5.70 Chlorite 2.19 Strontianite 0.53 Monazite 1.83 Biotite 2.40 Apatite 0.48 Kaolinite 0.41 Ilmenite 0.09 Rutile 6.55 Mica 1.56 SUM 98.33

Table A.5. Composition comparison between QEMSCAN, MLA, XRD, and XRF.

Composition Technique QEMSCAN (wt%) MLA (wt%) XRD (wt%) XRF (wt%) Size Fraction Technique 325 x 400 325 x 400 325 x 400 325 x 400 Barite 43.42 41.50 17.50 15.67 Calcite 16.23 14.20 24.20 32.69 Bastnaesite 10.29 12.80 5.20 10.14 Quartz 4.13 4.02 9.90 11.40 K_Feldspar 2.50 1.63 FeO 1.28 2.06 4.07 Parisite 2.02 2.74 Ankerite 5.30 2.16 Chlorite 1.94 1.23 Strontianite 0.75 1.31 3.31 Monazite 1.89 1.84 Biotite 1.64 0.67 Apatite 0.60 0.12 Kaolinite 0.30 0.17 Ilmenite 0.08 0.04 Rutile 4.93 0.28 Mica 1.27 1.13

Dolomite 9.92 43.00 7.66 SUM 98.57 97.82 99.80 84.94

Figure A.8. QEMSCAN mineral mass abundance in each size fraction.

APPENDIX B: SECOND LOT MINERALOGY

Table B.1. Microtrac particle size distributions for grind test samples

Grind time (Minutes) Grind time (Minutes) Size(um) 0 10 30 60 90 Size(um) 0 10 30 60 90 2000 100.00 100.00 100.00 100.00 100.00 31.11 8.84 7.27 23.76 38.69 55.02 1674 99.30 100.00 100.00 100.00 100.00 26.16 7.67 5.29 20.20 34.23 47.56 1408 98.15 100.00 100.00 100.00 100.00 22.00 6.61 3.48 16.89 30.29 41.40 1184 96.11 100.00 100.00 100.00 100.00 18.50 5.65 2.04 13.99 26.67 36.21 995.5 92.61 100.00 100.00 100.00 100.00 15.55 4.79 1.05 11.63 23.33 31.69 837.1 85.37 100.00 100.00 100.00 100.00 13.08 4.01 0.41 9.77 20.28 27.71 703.9 78.11 100.00 100.00 100.00 100.00 11.00 3.30 0.00 8.29 17.55 24.22 591.9 71.65 100.00 100.00 100.00 100.00 9.250 2.67 0.00 7.08 15.14 21.15 497.8 64.75 99.63 100.00 100.00 100.00 7.780 2.11 0.00 6.11 13.04 18.41 418.6 58.87 98.97 100.00 100.00 100.00 6.540 1.60 0.00 5.35 11.25 15.91 352.0 52.50 97.08 100.00 100.00 100.00 5.500 1.14 0.00 4.79 9.72 13.61 296.0 45.44 92.06 99.63 99.84 100.00 4.620 0.72 0.00 4.37 8.38 11.50 248.9 38.71 81.85 98.78 99.39 100.00 3.890 0.34 0.00 4.02 7.17 9.60 209.3 33.31 67.86 96.57 98.72 100.00 3.270 0.00 0.00 3.69 6.02 7.91 176.0 29.21 54.51 91.44 97.74 100.00 2.750 0.00 0.00 3.34 4.96 6.42 148.0 25.93 44.69 82.67 96.30 99.66 2.312 0.00 0.00 2.94 4.00 5.15 124.4 23.11 37.94 72.51 94.10 99.11 1.944 0.00 0.00 2.49 3.17 4.09 104.6 20.62 32.86 63.27 90.67 98.17 1.635 0.00 0.00 2.02 2.47 3.20 87.99 18.41 28.36 55.66 85.37 96.51 1.375 0.00 0.00 1.55 1.88 2.44 73.99 16.44 23.83 49.02 77.83 93.60 1.156 0.00 0.00 1.10 1.36 1.79 62.22 14.65 19.28 42.81 68.62 88.82 0.972 0.00 0.00 0.69 0.90 1.23 52.32 13.00 15.20 37.03 59.26 81.83 0.817 0.00 0.00 0.33 0.50 0.76 44.00 11.49 11.95 31.94 50.87 73.04 0.687 0.00 0.00 0.00 0.15 0.35 37.00 10.11 9.42 27.60 44.08 63.81 0.578 0.00 0.00 0.00 0.00 0.00

Table B.2. Modal mineral content of the bastnaesite ore (wt%).

10 30 60 90 Bastnaesite Minute Minute Minute Minute Mineral Formula Ore Grind Grind Grind Grind Barite BaSO4 24.6 26.4 26.1 27.3 30.9 Calcite CaCO3 21.3 20 20.2 19 17.5 Dolomite CaMg(CO3)2 11.6 11.2 10.9 10.9 9.24 Bastnaesite (Ce,La)(CO3)F 8.9 9.98 10.6 11.3 12.1 Quartz SiO2 7.65 7.95 7.66 7.38 6.64 K_Feldspar KAlSi3O8 6.4 5.84 6.01 5.81 5.78 Biotite K(Mg,Fe)3(AlSi3O10)(OH)2 3.2 2.29 20.9 2.03 1.81 Mica KAl2(AlSi3O10)(OH)2 2.97 2.68 2.48 2.21 2.12 FeO Fe3O4 2.36 2.14 2.27 2.53 2.35 Strontianite SrCO3 2.29 2.28 2.44 2.25 2.28 Parisite Ca(Ce,La)2(CO3)3F2 1.89 2.06 2.2 2.29 2.46 Ankerite CaFe(CO3)2 1.59 1.55 1.73 1.3 1.37 Celestine SrSO4 1.19 1.34 1.27 1.45 1.34 Monazite (La,Ce)PO4 0.99 1.31 1.27 1.57 1.62 Plagioclase (Na,Ca)(Al,Si)4O8 0.89 1.11 1.06 1.07 0.91 Chlorite (Mg3,Fe2)Al(AlSi3)O10(OH)8 0.76 0.72 0.65 0.57 0.46 Hornblende (Ca2,Na)(Mg2FeAl)Si6O22(OH)2 0.32 0.26 0.29 0.22 0.26 Allanite (Ca,Ce)2(Al,Fe)3(SiO4)(Si2O7)O(OH) 0.28 0.24 0.21 0.21 0.15 Kaolinite Al2Si2O5(OH)4 0.2 0.11 0.07 0.04 0.01 Apatite Ca5(PO4)3F 0.18 0.18 0.2 0.29 0.19 Galena PbS 0.16 0.1 0.11 0.11 0.13 Rutile TiO2 0.11 0.1 0.09 0.12 0.13 Ilmenite FeTiO3 0.08 0.05 0.05 0.03 0.04 Pyrite FeS2 0.07 0.03 0.02 0.02 0.03 Magnetoplumbite Pb(Fe,Mn)12O19 0.04 0.03 0.05 0.05 0.06 Pyroxene CaMgSi2O6 0.02 0.03 0.01 0.01 0.01 Zircon ZrSiO4 P 0.02 0.01 0.01 P Plumbophyllite Pb2Si4O10H2O P P P P P Thorite ThSiO4 P 0.01 P 0.01 0.02 Enargite_Tenn Cu3AsS4 P 0.03 ND PP

Figure B.1. Mineral locking for bastnaesite.

Figure B.2. Mineral locking for parisite.

Figure B.3. Mineral locking for monazite.

Figure B.4. Mineral locking for calcite.

Figure B.5. False-color image showing particle association of the +100 mesh size fraction.

Figure B.6. False-color image showing particle association of the +100 mesh size fraction.

Figure B.7. False-color image showing particle association of the +100 mesh size fraction.

Figure B.8. Measured and WPPF-calculated diffractograms and residual plot for the original bastnaesite ore.

Figure B.9. Bastnaesite ore diffractogram with candidate phases.

Table B.3. REE cumulative mass recovery to each liberation class

REE Cumulative Mass Recovery 10 30 60 90 Liberation Bastnaesite Minute Minute Minute Minute Class Ore Grind Grind Grind Grind 100 22 43 60 65 71 95-100 27 55 68 75 78 90-95 32 62 73 79 83 85-90 40 66 76 83 86 80-85 43 70 78 85 88 75-80 45 73 81 86 89 70-75 48 75 83 88 91 65-70 53 78 85 89 92 60-65 58 80 87 90 93 55-60 61 83 88 91 93 50-55 65 84 90 92 94 45-50 70 86 92 93 95 40-45 75 88 93 94 96 35-40 77 90 94 95 96 30-35 81 92 95 96 97 25-30 85 94 96 97 98 20-25 88 95 97 98 98 15-20 91 97 98 98 99 10-15 95 98 98 99 99 5-10 98 98 99 99 100 0-5 100 100 100 100 100

APPENDIX C: EXPERIMENTAL DATA

Table C.1. Experimental data for microflotation tests Feed (Calculated) Concentrate Tailings Depressant Collector Grade (%) Mass Grade (%) Recovery (%) Mass Grade (%) Recovery (%) Balance (%) Test Dose (M) Type Dose (M) Type Ca Ba TREE (g) Ca Ba TREE Mass Ca Ba TREE (g) Ca Ba TREE Mass Ca Ba TREE Mass Ca Ba TREE 9 1.00E-05 Sodium Oleate 10 33 3 0.48 9 45 3 48 40 66 51 0.44 12 20 3 44 52 26 41 92 92 92 92 1 5.00E-05 Sodium Oleate 13 22 3 0.76 13 22 3 76 77 79 79 0.11 12 16 2 11 10 8 8 87 87 87 87 11 1.00E-04 Sodium Oleate 11 30 3 0.69 11 34 3 69 68 78 73 0.21 12 17 2 21 22 12 17 90 90 90 90 16 1.00E-04 Hydroxamic Acid 14 19 3 0.68 13 23 3 68 64 86 67 0.19 16 1 4 19 23 1 20 87 87 87 87 12 3.00E-04 Hydroxamic Acid 13 24 3 0.84 13 24 4 84 87 81 87 0.07 8 34 2 7 4 10 4 91 91 91 91 10 5.00E-06 5.00E-05 Sodium Oleate 12 26 3 0.69 12 29 4 69 69 77 77 0.22 12 16 2 22 22 14 14 91 91 91 91 2 1.00E-05 5.00E-05 Sodium Oleate 11 33 3 0.63 10 39 3 63 59 74 69 0.30 12 21 2 30 34 19 24 93 93 93 93 Ammonium 5 2.00E-05 5.00E-05 Sodium Oleate 13 24 3 0.50 12 26 4 50 50 54 56 0.41 13 22 3 41 41 37 35 91 91 91 91 Lignin 17 5.00E-06 1.00E-04 Hydroxamic Acid 12 26 3 0.83 12 26 3 83 85 83 88 0.12 10 26 2 12 10 12 7 95 95 95 95 Sulfonate 19 1.00E-05 1.00E-04 Hydroxamic Acid 12 26 3 0.83 12 26 3 83 84 84 87 0.13 11 23 2 13 12 12 9 96 96 96 96 13 2.00E-05 1.00E-04 Hydroxamic Acid 11 31 3 0.74 11 32 3 74 75 77 78 0.12 10 24 2 12 11 9 8 86 86 86 86 6 100 5.00E-05 Sodium Oleate 12 29 3 0.70 12 34 3 70 71 81 69 0.22 11 15 3 22 21 11 23 92 92 92 92 3 300 5.00E-05 Sodium Oleate 12 29 3 0.71 12 31 3 71 70 76 68 0.13 12 17 4 13 14 8 16 84 84 84 84 Sodium 8 500 5.00E-05 Sodium Oleate 11 28 3 0.74 11 30 3 74 74 80 71 0.17 12 19 3 17 17 11 20 91 91 91 91 Silicate 18 100 1.00E-04 Hydroxamic Acid 11 33 3 0.70 11 35 3 70 70 74 71 0.25 11 27 3 25 25 21 24 95 95 95 95 (mg/L) 20 300 1.00E-04 Hydroxamic Acid 13 25 3 0.58 13 25 3 58 60 58 54 0.38 12 25 3 38 36 38 42 96 96 96 96 14 500 1.00E-04 Hydroxamic Acid 12 26 3 0.33 12 30 3 33 31 39 30 0.54 13 23 3 54 56 48 57 87 87 87 87 4 5.00E-07 5.00E-05 Sodium Oleate 12 26 3 0.65 13 28 4 65 66 70 71 0.16 11 18 2 16 15 11 10 81 81 81 81 7 1.00E-06Copper 5.00E-05 Sodium Oleate 11 29 3 0.53 11 35 3 53 49 65 57 0.41 12 20 3 41 45 29 37 94 94 94 94 21 5.00E-07Nitrate 1.00E-04 Hydroxamic Acid 11 31 3 0.75 11 30 3 75 78 73 81 0.18 9 35 2 18 15 20 12 93 93 93 93 15 1.00E-06 1.00E-04 Hydroxamic Acid 11 32 3 0.85 11 32 3 85 85 86 85 0.03 12 21 3 3 3 2 3 88 88 88 88

Table C.2. Experimental data for WHIMS tests Feed (Calculated) Magnetic Nonmagnetic

P80 Field Strength MassGrade (%) Mass Grade (%) Recovery (%) Mass Grade (%) Recovery (%) Balance (%) Standard Run (µm) (Gauss) (g) Ca Ba TREE (g) Ca Ba TREE Mass Ca Ba TREE (g) Ca Ba TREE Mass Ca Ba TREE Mass Ca Ba TREE 5 1 762 5000 50 12.7 6.7 2.1 4.4 10 4 2.4 8.8 6.9 5.2 9.9 44 13 7 2.1 88 90 92 87 97 97 97 97 4 2 50 5000 50 15.8 9.8 4.8 1.4 8 2 4.9 2.8 1.4 0.6 2.9 47.2 16 10 4.8 94 96 97 94 97 97 97 97 6 3 762 10000 50 11.0 5.5 2.2 8.4 11 3 3.7 16.8 16.8 9.2 28.2 41 11 6 1.9 82 82 90 71 99 99 99 99 1 4 144 7500 50 16.8 9.7 3.7 2 11 2 3.8 4.0 2.6 0.8 4.1 48.3 17 10 3.7 97 98 100 96 101 101 101 101 3 5 50 10000 50 15.9 8.5 4.9 3.3 14 2 8.4 6.6 5.8 1.5 11.4 46.2 16 9 4.6 92 93 97 88 99 99 99 99 2 6 144 7500 50 16.8 10.6 6.1 2.3 12 3 9.6 4.6 3.3 1.3 7.3 47 17 11 5.9 94 95 97 91 99 99 99 99

Table C.3. Experimental data for bench flotation tests Feed (Calculated) Concentrate Tailings Depressant Collector MassGrade (%) Mass Grade (%) Recovery (%) Mass Grade (%) Recovery (%) Balance (%) Test Dose (M) Type Dose (M) Type (g) Ca Ba TREE (g) Ca Ba TREE Mass Ca Ba TREE (g) Ca Ba TREE Mass Ca Ba TREE Mass Ca Ba TREE Ore Ammonium 500 14 11 6.1 78 9 10 30 16 10 14 77 405 15 11 1.5 81 87 82 9955 97 97 97 97 1.50E-03 1.00E-02 Hydroxamic Acid Gravity Con Lignin Sulfonate 500 12 15 8.9 127 5 12 29 25 10 20 82 347 15 16 1.6 69 84 74 6209 95 95 95 95

Table C.4. Experimental data for Heavy Liquid tests Feed (Calculated) Floats Sinks MassGrade (%) Mass Grade (%) Recovery (%) Mass Grade (%) Recovery (%)

Test P80 (µm) Specific Gravity (g) Ca Ba TREE (g) Ca Ba TREE Mass Ca Ba TREE (g) Ca Ba TREE Mass Ca Ba TREE 1 50 2.95 5.0 9 8 4.8 1.13 11 0.4 0.14 23 34 1.3 0.8 1.83 5 18 10.5 37 25 96 98 2 50 2.90 5.0 13 7 4.2 1.73 24 2 0.77 35 79 11 7.6 1.72 3 16 9.4 34 10 88 92 3 50 2.80 5.0 11 6 3.3 1.48 15 0.5 0.10 30 45 2.6 1.0 2.44 7 11 5.9 49 35 94 98 4 50 2.70 5.0 10 10 6.0 1.24 13 0.2 0.02 25 36 0.6 0.1 2.76 10 16 9.4 55 62 99 100 5 144 2.95 5.0 11 9 4.4 1.47 15 1.0 0.56 29 47 3.9 4.3 2.01 9 18 9.2 40 38 96 96 6 144 2.90 5.0 9 7 3.5 1.54 13 0.5 0.21 31 50 2.4 2.2 2.05 7 15 7.1 41 36 97 98 7 144 2.80 5.0 10 11 5.3 1.65 12 0.2 0.114 33 46 0.7 0.8 2.21 9 21 10.5 44 46 99 99 8 144 2.70 5.0 14 11 4.9 1.07 9 0 0.00 21 16 0.0 0.0 3.10 14 16 7.2 62 70 99 100 9 762 2.95 5.0 12 6 3.2 1.92 14 1.0 0.56 38 51 7.0 7.5 2.13 7 11 5.8 43 28 86 86 10 762 2.90 5.0 14 10 5.0 2.21 17 1.0 0.77 44 56 4.5 7.3 2.31 12 20 9.4 46 41 95 93 11 762 2.80 5.0 13 9 4.1 1.67 12 0.4 0.17 33 34 1.6 1.5 2.59 12 16 7.1 52 53 97 97 12 762 2.70 5.0 14 10 5.2 1.19 9 0.3 0.06 24 16 0.7 0.3 3.47 16 14 7 69 83 99 100

Middlings Mass Grade (%) Recovery (%) Balance (%)

Test P80 (µm) Specific Gravity (g) Ca Ba TREE Mass Ca Ba TREE Mass Ca Ba TREE 1 50 2.95 1.11 13 0.7 0.23 22 40 2.3 1.3 81 81 81 81 2 50 2.90 0.72 8 0.3 0.07 14 11 0.7 0.3 83 83 83 83 3 50 2.80 0.51 19 2.0 0.39 10 20 3.6 1.4 89 89 89 89 4 50 2.70 0.35 3 0.1 0.00 7 2 0.0 0.0 87 87 87 87 5 144 2.95 0.87 8 0.2 0.02 17 15 0.5 0.1 87 87 87 87 6 144 2.90 0.71 8 0.1 0.01 14 14 0.2 0.0 86 86 86 86 7 144 2.80 0.52 6 0.0 0.00 10 7 0.0 0.0 88 88 88 88 8 144 2.70 0.36 25 2.0 0.20 7 15 1.4 0.3 91 91 91 91 9 762 2.95 0.48 22 4.0 1.80 10 20 7.1 6.1 91 91 91 91 10 762 2.90 0.16 10 1.0 0.26 3 2 0.3 0.2 93 93 93 93 11 762 2.80 0.33 24 2.0 0.90 7 13 1.5 1.6 92 92 92 92 12 762 2.70 0.03 8 1.0 0.30 1 0.4 0.1 0.0 94 94 94 94

Table C.5. Experimental data for Deister table tests Feed (Calculated) Concentrate Tailings MassGrade (%) Mass Grade (%) Recovery (%) Mass Grade (%) Recovery (%) Test (g) Ca Ba TREE (g) Ca Ba TREE Mass Ca Ba TREE (g) Ca Ba TREE Mass Ca Ba TREE 1 Recoveries correspond to 500 g feed 500 15 11 3.8 26 8 21 14 5 3 10 19 127 14 14 3.6 25 24 32 24 2 Recoveries correspond to 500 g feed 500 15 11 3.5 53 12 16 7.8 11 8 16 23 180 14 14 3.4 36 33 47 35 3 Recoveries correspond to 500 g feed 500 15 11 3.8 63 9.3 20 9.2 13 8 24 31 160 14 14 3.4 32 30 43 29 4 Feed was Tailings from Test 3 (266 g) 266 16 8 2.4 25 13 14 5.4 9 7 17 21 32 15 10 3.2 12 11 17 16 5 Feed was Middlings from Test 3 (160 g) 160 14 12 3.5 6 17 7.3 2.5 4 4 2 3 120 14 13 3.5 75 74 78 76 6 Feed was Concentrate from Test 3 (63 g) 63 8 23 12.4 18 4.7 26 20 29 18 32 46 0 0 0 0.0 0 0 0 0 4* Recoveries correspond to initial 500g feed of Test 3 500 15 11 4.0 25 13 14 5.4 5 4 6 7 32 15 10 3.2 6 7 6 5 5* Recoveries correspond to initial 500g feed of Test 3 500 15 11 4.0 6 17 7 2.5 1 6 3 3 120 14 13 3.5 24 22 28 21 6* Recoveries correspond to initial 500g feed of Test 3 500 15 11 4.0 18 5 26 20 4 2 12 25 0 0 0 0.0 0 0 0 0

Middlings Mass Grade (%) Recovery (%) Balance (%) Test (g) Ca Ba TREE Mass Ca Ba TREE Mass Ca Ba TREE 1 Recoveries correspond to 500 g feed 314 16 9 3.0 63 67 51 50 93 93 93 93 2 Recoveries correspond to 500 g feed 265 17 7.4 2.8 53 58 37 41 100 100 100 100 3 Recoveries correspond to 500 g feed 266 17 6 2.7 53 60 31 38 98 98 98 98 4 Feed was Tailings from Test 3 (266 g) 201 17 6 1.9 76 78 63 60 97 97 97 97 5 Feed was Middlings from Test 3 (160 g) 24 15 11 3.4 15 16 13 15 94 94 94 94 6 Feed was Concentrate from Test 3 (63 g) 39 8.9 22 9.0 62 73 59 45 90 90 90 90 4* Recoveries correspond to initial 500g feed of Test 3 201 17 6 1.9 40 47 22 19 93 93 93 93 5* Recoveries correspond to initial 500g feed of Test 3 24 15 11 3.4 5 5 5 4 93 93 93 93 6* Recoveries correspond to initial 500g feed of Test 3 39 9 22 9.0 8 5 16 18 93 93 93 93

Table C.6. Experimental data for Falcon tests Feed (Calculated) Concentrate 1 Concentrate 2 Feed Rate MassGrade (%) Mass Grade (%) Recovery (%) Mass Grade (%) Recovery (%)

Run G-Force (G's) P80 (µm) (kg/hr) (g) Ca Ba TREE (g) Ca Ba TREE Mass Ca Ba TREE (g) Ca Ba TREE Mass Ca Ba TREE Opt 1 100 50 30 1278 14 8 5 87 16 16 9.5 7 8 13 14 86 15 14 8.3 7 7 12 12 Opt 2 100 50 30 1199 14 8 4 93 14 14 9.3 8 8 14 19 85 16 14 8.2 7 8 13 15 Opt 3 100 50 30 1214 12 7 3 101 11 11 7.2 8 7 13 18 94 14 13 7.1 8 9 14 16 1 100 50 30 1213 14 9 5 89 12 16 10.6 7 6 13 17 80.9 11 13 7.3 7 5 10 10 2 250 144 30 1194 17 10 4 130 15 11 6 11 10 12 15 68.1 16 11 5.9 6 5 7 8 3 250 50 60 1097 13 7 4 128 13 12 7.1 12 12 19 23 89.1 16 14 7.2 8 10 15 16 4 100 50 60 1150 16 8 4 121 14 11 6.1 11 9 14 18 134.6 17 12 5.9 12 13 18 19 5 175 77 45 1160 15 9 4 119 15 13 7.2 10 10 15 17 91.5 17 14 7.2 8 9 13 13 6 250 144 60 1272 15 8 4 103 13 11 6 8 7 11 11 109.8 16 13 7.1 9 9 14 14 7 175 77 45 1179 9 5 2 102 10 8 5.8 9 9 15 21 88.4 13 10 6 7 10 16 19 8 250 50 30 1142 17 9 5 92 17 14 9.2 8 8 12 16 89.5 18 13 7.2 8 8 11 12 9 100 144 30 1270 17 9 4 192 17 13 7.1 15 15 21 25 111.6 13 9 4.8 9 7 9 10 10 100 144 60 1177 16 9 4 110 15 14 7.3 9 9 14 15 104.6 16 13 7.1 9 9 12 14 11 175 77 45 1230 14 7 4 100 13 10 6 8 8 11 12 94.2 10 7 3.7 8 6 7 7

Concentrate 3 Tailings Feed Rate Mass Grade (%) Recovery (%) Mass Grade (%) Recovery (%) Balance (%)

Run G-Force (G's) P80 (µm) (kg/hr) (g) Ca Ba TREE Mass Ca Ba TREE (g) Ca Ba TREE Mass Ca Ba TREE Mass Ca Ba TREE Opt 1 81 16 14 8.2 6 7 11 11 806 13 6 3.3 63 60 47 45 83 83 83 83 Opt 2 67 15 11 5.9 6 6 8 9 715 14 6 2.4 60 59 46 37 80 80 80 80 Opt 3 71 16 12 5.9 6 8 10 10 605 12 5 1.9 50 48 35 28 72 72 72 72 1 100 90 30 71.2 15 14 7.2 6 6 9 9 794.1 15 7 3.5 65 68 53 49 85 85 85 85 2 250 30 30 44.5 19 9 3.6 4 4 3 3 389.7 17 9 3.6 33 33 31 27 53 53 53 53 3 250 90 60 75.8 15 9 3.8 7 8 8 7 556.4 12 5 2.2 51 47 34 31 77 77 77 77 4 100 90 60 123.5 16 11 5.8 11 11 15 18 624.6 16 6 2.1 54 55 41 32 87 87 87 87 5 175 60 45 96.7 14 9 3.7 8 8 9 7 605.6 15 7 3.5 52 52 42 42 79 79 79 79 6 250 30 60 105.3 16 10 5.8 8 9 11 11 633.3 15 6 3.4 50 50 38 39 75 75 75 75 7 175 60 45 85.4 16 11 5.9 7 12 17 18 746.3 8 3 1.08 63 54 40 29 87 87 87 87 8 250 90 30 86.7 14 10 5.8 8 6 8 10 698.9 17 8 3.4 61 62 53 46 85 85 85 85 9 100 30 30 86 17 10 5.8 7 7 7 9 651.1 18 8 3.3 51 54 45 39 82 82 82 82 10 100 30 60 94.5 15 9 3.7 8 8 8 7 603.8 16 8 3.6 51 52 44 41 78 78 78 78 11 175 60 45 88.1 15 9 5.8 7 8 9 11 789.1 14 7 3.5 64 66 60 57 87 87 87 87

100

APPENDIX D: ECONOMIC ESTIMATES

Figure D.1. Sensitivity Analysis for the flotation concentrator.

101

Table D.1. Cash flow diagram for the flotation concentrator

Operating Year Initial 1 2 3 4 5 6 7 8 9 10

Revenue REO Price x REO Recovered $/year$ 83,200,000 $ 83,200,000 $ 83,200,000 $ 83,200,000 $ 83,200,000 $ 83,200,000 $ 83,200,000 $ 83,200,000 $ 83,200,000 $ 83,200,000 Ore Mined 400,000 tonnes/year 400,000 400,000 400,000 400,000 400,000 400,000 400,000 400,000 400,000 400,000 REO Mined Ore Mined x REO Grade tonnes/year 32,000 32,000 32,000 32,000 32,000 32,000 32,000 32,000 32,000 32,000 REO Recovered REO Mined x REO Recovery tonnes/year 20,800 20,800 20,800 20,800 20,800 20,800 20,800 20,800 20,800 20,800

OPEX Sum of Operating Costs: $/year$ 13,809,728 $ 13,809,728 $ 13,809,728 $ 13,809,728 $ 13,809,728 $ 13,809,728 $ 13,809,728 $ 13,809,728 $ 13,809,728 $ 13,809,728 Labor Labor Rate x Schedule $/year 3,485,720 3,485,720 3,485,720 3,485,720 3,485,720 3,485,720 3,485,720 3,485,720 3,485,720 3,485,720 Reagents Reagent Cost x Dose x Ore Mined $/year 9,403,325 9,403,325 9,403,325 9,403,325 9,403,325 9,403,325 9,403,325 9,403,325 9,403,325 9,403,325 Energy Energy Cost x Ore Mined $/year 920,683 920,683 920,683 920,683 920,683 920,683 920,683 920,683 920,683 920,683

Profit Revenue - OPEX $/year$ 69,390,272 $ 69,390,272 $ 69,390,272 $ 69,390,272 $ 69,390,272 $ 69,390,272 $ 69,390,272 $ 69,390,272 $ 69,390,272 $ 69,390,272

Investment $/year ($19,318,444) ($166,667) ($166,667) ($166,667) ($166,667) ($166,667) ($166,667) ($166,667) ($166,667) ($166,667) ($166,667) Initial Capital Equipment $/year $ (19,318,444) Sustaining $500k / 3 years $/year ($166,667) ($166,667) ($166,667) ($166,667) ($166,667) ($166,667) ($166,667) ($166,667) ($166,667) ($166,667)

Cash Investment + Profit $/year $ (19,318,444) $ 69,223,605 $ 69,223,605 $ 69,223,605 $ 69,223,605 $ 69,223,605 $ 69,223,605 $ 69,223,605 $ 69,223,605 $ 69,223,605 $ 69,223,605 Cash Flow $/year $ (19,318,444) $ 49,905,161 $ 119,128,766 $ 188,352,372 $ 257,575,977 $ 326,799,583 $ 396,023,188 $ 465,246,793 $ 534,470,399 $ 603,694,004 $ 672,917,609

NPV $ (19,318,444) $ 46,208,483 $ 102,133,716 $ 149,520,185 $ 189,326,033 $ 222,414,305 $ 249,561,785 $ 271,467,035 $ 288,757,726 $ 301,997,302 $ 311,691,055

NPV - 20% $ (19,318,444) $ 6,149,223 $ 27,949,903 $ 46,487,111 $ 62,124,706 $ 75,190,547 $ 85,979,832 $ 94,758,136 $ 101,764,182 $ 107,212,360 $ 111,295,024 NPV - 15% $ (19,318,444) $ 8,460,334 $ 32,229,738 $ 52,431,327 $ 69,463,244 $ 83,684,226 $ 95,417,252 $ 104,952,880 $ 112,552,271 $ 118,449,953 $ 122,856,333 NPV - 10% $ (19,318,444) $ 10,771,446 $ 36,509,573 $ 58,375,543 $ 76,801,782 $ 92,177,904 $ 104,854,673 $ 115,147,624 $ 123,340,360 $ 129,687,546 $ 134,417,643 NPV - 5% $ (19,318,444) $ 13,082,557 $ 40,789,409 $ 64,319,759 $ 84,140,321 $ 100,671,582 $ 114,292,093 $ 125,342,369 $ 134,128,449 $ 140,925,138 $ 145,978,952 NPV - No Change $ (19,318,444) $ 15,393,668 $ 45,069,244 $ 70,263,974 $ 91,478,859 $ 109,165,261 $ 123,729,513 $ 135,537,113 $ 144,916,538 $ 152,162,731 $ 157,540,262

REOPrice NPV + 5% $ (19,318,444) $ 17,704,779 $ 49,349,080 $ 76,208,190 $ 98,817,397 $ 117,658,939 $ 133,166,934 $ 145,731,857 $ 155,704,627 $ 163,400,324 $ 169,101,571 NPV +10% $ (19,318,444) $ 20,015,890 $ 53,628,915 $ 82,152,406 $ 106,155,935 $ 126,152,617 $ 142,604,354 $ 155,926,601 $ 166,492,716 $ 174,637,917 $ 180,662,881 NPV +15% $ (19,318,444) $ 22,327,001 $ 57,908,750 $ 88,096,622 $ 113,494,473 $ 134,646,296 $ 152,041,774 $ 166,121,345 $ 177,280,805 $ 185,875,510 $ 192,224,190 NPV +20% $ (19,318,444) $ 24,638,112 $ 62,188,586 $ 94,040,838 $ 120,833,011 $ 143,139,974 $ 161,479,195 $ 176,316,090 $ 188,068,895 $ 197,113,102 $ 203,785,500 NPV - 20% $ (19,318,444) $ 17,135,024 $ 48,293,978 $ 74,742,772 $ 97,008,238 $ 115,565,005 $ 130,840,341 $ 143,218,562 $ 153,045,056 $ 160,629,938 $ 166,251,379 NPV - 15% $ (19,318,444) $ 16,699,685 $ 47,487,795 $ 73,623,073 $ 95,625,893 $ 113,965,069 $ 129,062,634 $ 141,298,200 $ 151,012,927 $ 158,513,136 $ 164,073,600 NPV - 10% $ (19,318,444) $ 16,264,346 $ 46,681,611 $ 72,503,373 $ 94,243,548 $ 112,365,133 $ 127,284,927 $ 139,377,838 $ 148,980,797 $ 156,396,334 $ 161,895,821 NPV - 5% $ (19,318,444) $ 15,829,007 $ 45,875,428 $ 71,383,674 $ 92,861,203 $ 110,765,197 $ 125,507,220 $ 137,457,475 $ 146,948,668 $ 154,279,533 $ 159,718,041 NPV - No Change $ (19,318,444) $ 15,393,668 $ 45,069,244 $ 70,263,974 $ 91,478,859 $ 109,165,261 $ 123,729,513 $ 135,537,113 $ 144,916,538 $ 152,162,731 $ 157,540,262 NPV + 5% $ (19,318,444) $ 14,958,329 $ 44,263,061 $ 69,144,275 $ 90,096,514 $ 107,565,324 $ 121,951,807 $ 133,616,750 $ 142,884,409 $ 150,045,930 $ 155,362,482 NPV +10% $ (19,318,444) $ 14,522,990 $ 43,456,877 $ 68,024,576 $ 88,714,169 $ 105,965,388 $ 120,174,100 $ 131,696,388 $ 140,852,279 $ 147,929,128 $ 153,184,703 OPEX(Reagents) NPV +15% $ (19,318,444) $ 14,087,650 $ 42,650,694 $ 66,904,876 $ 87,331,824 $ 104,365,452 $ 118,396,393 $ 129,776,026 $ 138,820,150 $ 145,812,327 $ 151,006,924 NPV +20% $ (19,318,444) $ 13,652,311 $ 41,844,510 $ 65,785,177 $ 85,949,479 $ 102,765,516 $ 116,618,686 $ 127,855,663 $ 136,788,020 $ 143,695,525 $ 148,829,144 NPV - 20% $ (19,318,444) $ 10,481,196 $ 33,896,628 $ 53,785,597 $ 70,542,062 $ 84,520,969 $ 96,041,864 $ 105,392,196 $ 112,830,314 $ 118,588,206 $ 122,873,980 NPV - 15% $ (19,318,444) $ 11,681,158 $ 36,655,443 $ 57,865,740 $ 75,732,643 $ 90,635,090 $ 102,914,225 $ 112,876,916 $ 120,798,967 $ 126,928,028 $ 131,486,259 NPV - 10% $ (19,318,444) $ 12,900,829 $ 39,438,267 $ 61,973,446 $ 80,953,684 $ 96,781,988 $ 109,821,168 $ 120,397,578 $ 128,804,527 $ 135,305,384 $ 140,136,405 NPV - 5% $ (19,318,444) $ 14,138,757 $ 42,243,373 $ 66,106,764 $ 86,203,049 $ 102,959,382 $ 116,760,301 $ 127,951,705 $ 136,844,463 $ 143,717,708 $ 148,821,833 NPV - No Change $ (19,318,444) $ 15,393,668 $ 45,069,244 $ 70,263,974 $ 91,478,859 $ 109,165,261 $ 123,729,513 $ 135,537,113 $ 144,916,538 $ 152,162,731 $ 157,540,262 NPV + 5% $ (19,318,444) $ 16,664,436 $ 47,914,537 $ 74,443,555 $ 96,779,444 $ 115,397,842 $ 130,726,933 $ 143,151,863 $ 153,018,769 $ 160,638,441 $ 166,289,663 NPV +10% $ (19,318,444) $ 17,950,061 $ 50,778,054 $ 78,644,149 $ 102,103,316 $ 121,655,532 $ 137,750,887 $ 150,794,224 $ 161,149,381 $ 169,143,037 $ 175,068,226

Production (tonnes/day Production NPV +15% $ (19,318,444) $ 19,249,648 $ 53,658,724 $ 82,864,537 $ 107,449,140 $ 127,936,904 $ 144,799,874 $ 158,462,640 $ 169,306,780 $ 177,674,902 $ 183,874,320 NPV +20% $ (19,318,444) $ 20,562,390 $ 56,555,583 $ 87,103,625 $ 112,815,712 $ 134,240,667 $ 151,872,538 $ 166,155,706 $ 177,489,529 $ 186,232,576 $ 192,706,477 NPV - 20% $ (19,318,444) $ 18,971,157 $ 48,381,735 $ 73,331,095 $ 94,318,785 $ 111,794,822 $ 126,164,293 $ 137,791,538 $ 147,003,969 $ 154,095,538 $ 159,329,897 NPV - 15% $ (19,318,444) $ 18,076,785 $ 47,553,612 $ 72,564,315 $ 93,608,804 $ 111,137,432 $ 125,555,598 $ 137,227,932 $ 146,482,111 $ 153,612,336 $ 158,882,488 NPV - 10% $ (19,318,444) $ 17,182,413 $ 46,725,489 $ 71,797,535 $ 92,898,822 $ 110,480,042 $ 124,946,903 $ 136,664,325 $ 145,960,254 $ 153,129,134 $ 158,435,080 NPV - 5% $ (19,318,444) $ 16,288,040 $ 45,897,367 $ 71,030,755 $ 92,188,840 $ 109,822,651 $ 124,338,208 $ 136,100,719 $ 145,438,396 $ 152,645,933 $ 157,987,671 NPV - No Change $ (19,318,444) $ 15,393,668 $ 45,069,244 $ 70,263,974 $ 91,478,859 $ 109,165,261 $ 123,729,513 $ 135,537,113 $ 144,916,538 $ 152,162,731 $ 157,540,262 CAPEX NPV + 5% $ (19,318,444) $ 14,499,295 $ 44,241,122 $ 69,497,194 $ 90,768,877 $ 108,507,870 $ 123,120,819 $ 134,973,506 $ 144,394,680 $ 151,679,530 $ 157,092,853 NPV +10% $ (19,318,444) $ 13,604,923 $ 43,412,999 $ 68,730,414 $ 90,058,895 $ 107,850,480 $ 122,512,124 $ 134,409,900 $ 143,872,823 $ 151,196,328 $ 156,645,444 NPV +15% $ (19,318,444) $ 12,710,550 $ 42,584,876 $ 67,963,634 $ 89,348,914 $ 107,193,089 $ 121,903,429 $ 133,846,294 $ 143,350,965 $ 150,713,127 $ 156,198,035 NPV +20% $ (19,318,444) $ 11,816,178 $ 41,756,754 $ 67,196,854 $ 88,638,932 $ 106,535,699 $ 121,294,734 $ 133,282,687 $ 142,829,107 $ 150,229,925 $ 155,750,626 Rate 0.08

Payback Period months$ 3.35 Internal Rate of Return 358%

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Table D.2. Cash flow diagram for the gravity-flotation concentrator

Operating Year Initial 1 2 3 4 5 6 7 8 9 10

Revenue REO Price x REO Recovered $/year$ 104,960,000 $ 104,960,000 $ 104,960,000 $ 104,960,000 $ 104,960,000 $ 104,960,000 $ 104,960,000 $ 104,960,000 $ 104,960,000 $ 104,960,000 Ore Mined 400,000 tonnes/year 400,000 400,000 400,000 400,000 400,000 400,000 400,000 400,000 400,000 400,000 REO Mined Ore Mined x REO Grade tonnes/year 32,000 32,000 32,000 32,000 32,000 32,000 32,000 32,000 32,000 32,000 REO Recovered REO Mined x REO Recovery tonnes/year 26,240 26,240 26,240 26,240 26,240 26,240 26,240 26,240 26,240 26,240

OPEX Sum of Operating Costs: $/year$ 13,205,885 $ 13,205,887 $ 13,205,889 $ 13,205,891 $ 13,205,893 $ 13,205,895 $ 13,205,897 $ 13,205,899 $ 13,205,901 $ 13,205,903 Labor Labor Rate x Schedule $/year 3,770,420 3,770,421 3,770,422 3,770,423 3,770,424 3,770,425 3,770,426 3,770,427 3,770,428 3,770,429 Reagents Reagent Cost x Dose x Ore Mined $/year 8,496,108 8,496,108 8,496,108 8,496,108 8,496,108 8,496,108 8,496,108 8,496,108 8,496,108 8,496,108 Energy Energy Cost x Ore Mined $/year 939,357 939,358 939,359 939,360 939,361 939,362 939,363 939,364 939,365 939,366

Profit Revenue - OPEX $/year$ 91,754,115 $ 91,754,113 $ 91,754,111 $ 91,754,109 $ 91,754,107 $ 91,754,105 $ 91,754,103 $ 91,754,101 $ 91,754,099 $ 91,754,097

Investment $/year ($22,254,940) ($166,667) ($166,667) ($166,667) ($166,667) ($166,667) ($166,667) ($166,667) ($166,667) ($166,667) ($166,667) Initial Capital Equipment $/year $ (22,254,940) Sustaining $500k / 3 years $/year ($166,667) ($166,667) ($166,667) ($166,667) ($166,667) ($166,667) ($166,667) ($166,667) ($166,667) ($166,667)

Cash Investment + Profit $/year $ (22,254,940) $ 91,587,448 $ 91,587,446 $ 91,587,444 $ 91,587,442 $ 91,587,440 $ 91,587,438 $ 91,587,436 $ 91,587,434 $ 91,587,432 $ 91,587,430 Cash Flow $/year $ (22,254,940) $ 69,332,509 $ 160,919,955 $ 252,507,400 $ 344,094,842 $ 435,682,282 $ 527,269,721 $ 618,857,157 $ 710,444,591 $ 802,032,024 $ 893,619,454

NPV $ (22,254,940) $ 64,196,767 $ 137,962,925 $ 200,448,515 $ 252,919,981 $ 296,518,041 $ 332,269,363 $ 361,097,207 $ 383,831,107 $ 401,215,692 $ 413,918,712

NPV - 20% $ (22,254,940) $ 13,660,472 $ 44,377,193 $ 70,468,335 $ 92,450,626 $ 110,789,624 $ 125,904,459 $ 138,172,161 $ 147,931,591 $ 155,487,034 $ 161,111,456 NPV - 15% $ (22,254,940) $ 16,576,027 $ 49,776,370 $ 77,967,192 $ 101,708,474 $ 121,504,725 $ 137,810,128 $ 151,033,223 $ 161,541,180 $ 169,663,689 $ 175,696,492 NPV - 10% $ (22,254,940) $ 19,491,583 $ 55,175,547 $ 85,466,049 $ 110,966,322 $ 132,219,827 $ 149,715,797 $ 163,894,284 $ 175,150,770 $ 183,840,345 $ 190,281,529 NPV - 5% $ (22,254,940) $ 22,407,138 $ 60,574,724 $ 92,964,905 $ 120,224,170 $ 142,934,929 $ 161,621,465 $ 176,755,346 $ 188,760,359 $ 198,017,000 $ 204,866,565 NPV - No Change $ (22,254,940) $ 25,322,694 $ 65,973,901 $ 100,463,762 $ 129,482,018 $ 153,650,031 $ 173,527,134 $ 189,616,408 $ 202,369,948 $ 212,193,656 $ 219,451,602

REOPrice NPV + 5% $ (22,254,940) $ 28,238,249 $ 71,373,078 $ 107,962,619 $ 138,739,866 $ 164,365,133 $ 185,432,803 $ 202,477,470 $ 215,979,538 $ 226,370,311 $ 234,036,638 NPV +10% $ (22,254,940) $ 31,153,805 $ 76,772,255 $ 115,461,476 $ 147,997,714 $ 175,080,235 $ 197,338,472 $ 215,338,532 $ 229,589,127 $ 240,546,967 $ 248,621,675 NPV +15% $ (22,254,940) $ 34,069,360 $ 82,171,432 $ 122,960,333 $ 157,255,562 $ 185,795,337 $ 209,244,140 $ 228,199,594 $ 243,198,716 $ 254,723,622 $ 263,206,711 NPV +20% $ (22,254,940) $ 36,984,916 $ 87,570,609 $ 130,459,190 $ 166,513,410 $ 196,510,438 $ 221,149,809 $ 241,060,656 $ 256,808,306 $ 268,900,278 $ 277,791,748 NPV - 20% $ (22,254,940) $ 26,896,047 $ 68,887,518 $ 104,510,453 $ 134,477,932 $ 159,432,339 $ 179,951,921 $ 196,556,765 $ 209,714,241 $ 219,843,960 $ 227,322,286 NPV - 15% $ (22,254,940) $ 26,502,709 $ 68,159,114 $ 103,498,780 $ 133,228,954 $ 157,986,762 $ 178,345,724 $ 194,821,676 $ 207,878,168 $ 217,931,384 $ 225,354,615 NPV - 10% $ (22,254,940) $ 26,109,370 $ 67,430,710 $ 102,487,108 $ 131,979,975 $ 156,541,185 $ 176,739,528 $ 193,086,586 $ 206,042,095 $ 216,018,808 $ 223,386,944 NPV - 5% $ (22,254,940) $ 25,716,032 $ 66,702,305 $ 101,475,435 $ 130,730,996 $ 155,095,608 $ 175,133,331 $ 191,351,497 $ 204,206,021 $ 214,106,232 $ 221,419,273 NPV - No Change $ (22,254,940) $ 25,322,694 $ 65,973,901 $ 100,463,762 $ 129,482,018 $ 153,650,031 $ 173,527,134 $ 189,616,408 $ 202,369,948 $ 212,193,656 $ 219,451,602 NPV + 5% $ (22,254,940) $ 24,929,355 $ 65,245,497 $ 99,452,090 $ 128,233,039 $ 152,204,454 $ 171,920,937 $ 187,881,319 $ 200,533,875 $ 210,281,080 $ 217,483,931 NPV +10% $ (22,254,940) $ 24,536,017 $ 64,517,092 $ 98,440,417 $ 126,984,061 $ 150,758,877 $ 170,314,741 $ 186,146,230 $ 198,697,802 $ 208,368,504 $ 215,516,260 OPEX(Reagents) NPV +15% $ (22,254,940) $ 24,142,679 $ 63,788,688 $ 97,428,744 $ 125,735,082 $ 149,313,300 $ 168,708,544 $ 184,411,141 $ 196,861,729 $ 206,455,927 $ 213,548,589 NPV +20% $ (22,254,940) $ 23,749,340 $ 63,060,284 $ 96,417,072 $ 124,486,103 $ 147,867,723 $ 167,102,347 $ 182,676,052 $ 195,025,656 $ 204,543,351 $ 211,580,918 NPV - 20% $ (22,254,940) $ 18,241,752 $ 50,470,121 $ 77,823,826 $ 100,848,201 $ 120,034,542 $ 135,825,110 $ 148,617,710 $ 158,769,841 $ 166,602,492 $ 172,403,585 NPV - 15% $ (22,254,940) $ 19,981,300 $ 54,309,745 $ 83,442,857 $ 107,961,957 $ 128,390,750 $ 145,200,671 $ 158,815,757 $ 169,617,082 $ 177,946,794 $ 184,111,790 NPV - 10% $ (22,254,940) $ 21,742,334 $ 58,174,771 $ 89,090,511 $ 115,106,937 $ 136,780,242 $ 154,611,098 $ 169,049,838 $ 180,501,158 $ 189,328,415 $ 195,857,527 NPV - 5% $ (22,254,940) $ 23,523,264 $ 62,063,362 $ 94,764,745 $ 122,280,938 $ 145,200,687 $ 164,053,966 $ 179,317,455 $ 191,419,528 $ 200,744,791 $ 207,638,222 NPV - No Change $ (22,254,940) $ 25,322,694 $ 65,973,901 $ 100,463,762 $ 129,482,018 $ 153,650,031 $ 173,527,134 $ 189,616,408 $ 202,369,948 $ 212,193,656 $ 219,451,602 NPV + 5% $ (22,254,940) $ 27,139,390 $ 69,904,959 $ 106,185,976 $ 136,708,461 $ 162,126,457 $ 183,028,709 $ 199,944,749 $ 213,350,436 $ 223,673,004 $ 231,295,656 NPV +10% $ (22,254,940) $ 28,972,255 $ 73,855,263 $ 111,929,966 $ 143,958,731 $ 170,628,337 $ 192,556,994 $ 210,300,728 $ 224,359,207 $ 235,181,035 $ 243,168,577

Production (tonnes/day Production NPV +15% $ (22,254,940) $ 30,820,311 $ 77,823,672 $ 117,694,462 $ 151,231,452 $ 179,154,214 $ 202,110,468 $ 220,682,781 $ 235,394,668 $ 246,716,136 $ 255,068,746 NPV +20% $ (22,254,940) $ 32,682,677 $ 81,809,163 $ 123,478,322 $ 158,525,387 $ 187,702,774 $ 211,687,761 $ 231,089,497 $ 246,455,378 $ 258,276,851 $ 266,994,702 NPV - 20% $ (22,254,940) $ 29,443,979 $ 69,789,906 $ 103,997,100 $ 132,753,627 $ 156,679,299 $ 176,332,011 $ 192,213,517 $ 204,774,679 $ 214,420,258 $ 221,513,270 NPV - 15% $ (22,254,940) $ 28,413,658 $ 68,835,905 $ 103,113,766 $ 131,935,725 $ 155,921,982 $ 175,630,792 $ 191,564,240 $ 204,173,496 $ 213,863,607 $ 220,997,853 NPV - 10% $ (22,254,940) $ 27,383,336 $ 67,881,903 $ 102,230,431 $ 131,117,822 $ 155,164,665 $ 174,929,573 $ 190,914,963 $ 203,572,314 $ 213,306,957 $ 220,482,436 NPV - 5% $ (22,254,940) $ 26,353,015 $ 66,927,902 $ 101,347,097 $ 130,299,920 $ 154,407,348 $ 174,228,353 $ 190,265,685 $ 202,971,131 $ 212,750,306 $ 219,967,019 NPV - No Change $ (22,254,940) $ 25,322,694 $ 65,973,901 $ 100,463,762 $ 129,482,018 $ 153,650,031 $ 173,527,134 $ 189,616,408 $ 202,369,948 $ 212,193,656 $ 219,451,602 CAPEX NPV + 5% $ (22,254,940) $ 24,292,372 $ 65,019,900 $ 99,580,428 $ 128,664,116 $ 152,892,714 $ 172,825,915 $ 188,967,131 $ 201,768,766 $ 211,637,005 $ 218,936,185 NPV +10% $ (22,254,940) $ 23,262,051 $ 64,065,899 $ 98,697,093 $ 127,846,213 $ 152,135,397 $ 172,124,695 $ 188,317,854 $ 201,167,583 $ 211,080,355 $ 218,420,767 NPV +15% $ (22,254,940) $ 22,231,730 $ 63,111,897 $ 97,813,759 $ 127,028,311 $ 151,378,080 $ 171,423,476 $ 187,668,577 $ 200,566,401 $ 210,523,704 $ 217,905,350 NPV +20% $ (22,254,940) $ 21,201,409 $ 62,157,896 $ 96,930,425 $ 126,210,409 $ 150,620,763 $ 170,722,257 $ 187,019,300 $ 199,965,218 $ 209,967,054 $ 217,389,933 Rate 0.08

Payback Period months$ 2.92 Internal Rate of Return 412%

103