Using Flotation Response to Infer an Operational Contact Angle of Particles

Transcription

1 Using Flotation Response to Infer an Operational Contact Angle of Particles by Solomon Muganda BSc. Engineering Hons. (Metallurgy) MPhil in Engineering (Metallurgy) Ian Wark Research Institute University of South Australia A thesis submitted for the degree of Doctor of Philosophy in Engineering (Minerals and Materials) February, 2010

2 Disclaimer This is to certify that this thesis contains no material which has been accepted for the award of any other degree or diploma, and that, to the best of the candidate s knowledge and belief, this thesis contains no material previously published or written by another person, except when due reference is made in the text of the thesis.... Solomon Muganda February, 2010

3 Acknowledgements I would like to thank Professor Stephen Grano and Dr Massimiliano Zanin for their supervision and guidance throughout my candidature. Their compliments and criticisms are well appreciated and this thesis is proof of the academic experience at the Ian Wark Research Institute. Many thanks go to Mr Czeslaw Poprawski for the tremendous laboratory assistance he offered. Laboratory training on equipment use and health and safety issues by Mr Keith Quast and Dr Ray Newell is appreciated. I also want to thank the following people for the time spent training me in their areas of competence: Dr Igor Ametov (laboratory flotation procedures), Dr Sabina Gredelj (bubble size measurement), Dr Denis Palms (Washburn contact angle measurement using the Wilhelmy balance). Discussions and collaborative work in areas of common research interest with fellow students are appreciated in the case of the following: Bian Xun (contact angle measurement with the Equilibrium Capillary Pressure technique), Susana (Contact angle/tof-sims surface analysis), and Daniel Chipfunhu (bubble size measurement). To Christopher Bassell, thanks for carrying out surface analysis for my samples with XPS. To my family and friends, I thank you for your support and goodwill towards my studies. God bless you all. A big thank you to the scholarship providers: the Ian Wark Research Institute and University of South Australia. Finally, to my wife, Neatness, I dedicate this thesis and thank you for the support and encouragement to complete this programme. i

4 Publications Muganda, S., Zanin, M. & Grano, S. (2008). Flotation behaviour of sulphide mineral size fractions with controlled contact angle. Chemeca Conference Proceedings: Towards a sustainable Australasia. Newcastle City Hall, New South Wales, Australia, Engineers Australia. Muganda, S., Zanin, M. & Grano, S. R. (2010). Influence of particle size and contact angle on the flotation of chalcopyrite in a laboratory batch flotation cell. International Journal of Mineral Processing, Submitted for Publication. Muganda, S., Zanin, M. & Grano, S. R. (2010). Inferring an operational contact angle on particles by flotation: Part 1 - single minerals. International Journal of Mineral Processing, Submitted for publication. Muganda, S., Zanin, M., & Grano, S. R. (2010). Inferring an operational contact angle on particles by flotation: Part 2-natural ore. International Journal of Mineral Processing, Submitted for publication. ii

5 Abstract The contact angle is a measure of the hydrophobicity of a surface and profoundly influences the wetting characteristics and flotation behaviour of minerals. Current technologies to quantify the contact angle of value mineral particles in a natural ore are very limited. This thesis principally addresses the need to determine the contact angle of value mineral particles within natural ores in which the value mineral occurs at low concentration, and in which issues of mineral locking and surface alteration (by for example surface oxidation) may cause the value mineral contact angle to vary with particle size fraction. Current methods to measure the contact angle, for example, by wetting in particle beds, are not sensitive enough to probe the contact angle on value minerals within samples of low concentration, a comment which is validated in this thesis. Moreover, variations in contact angle with particle size fraction also necessitate the development of new approaches that can measure, or infer, the contact angle as a function of particle size fraction. This thesis considers an alternative approach by way of benchmarking natural ore flotation behaviour against pre-determined calibration or master curves of floatability against both particle contact angle range and size fraction. In this thesis, calibration curves were determined from flotation data obtained on model chalcopyrite particles under well-defined and closely controlled physical conditions. The chalcopyrite was as a single, fully liberated mineral, thus, differences in floatability within individual particle size fractions were interpreted in terms of differences in particle hydrophobicity. This interpretation was confirmed by independent measurements of contact angle on the flotation feed and product samples. This alternative approach necessarily required testing the determined calibration curves by other independent measurements of contact angle on the ore particles of interest. This particular problem is tackled in the current thesis. The approach also required very close control of the flotation conditions to obtain acceptable reproducibility within defined error limits. Furthermore, validation that the particle size fractions floated independently of each other allowed the development of the master curves using data collated from many different flotation tests with different particle contact angle ranges within each particle size fraction. Further, benchmarking the ore flotation behaviour against the calibration curves iii

6 required that the effect of pulp density of the diluting gangue mineral be addressed. The applicability of the master curves, which were determined exclusively using the chalcopyrite-amyl xanthate collector system, was evaluated by testing other mineral collector systems, notably the chalcopyrite-dithiophosphate, pyrite-xanthate and pyritedithiophosphate systems. Once validated, the master curves were then used to benchmark ore flotation behaviour to infer an operational contact angle as a function of particle size fraction for specific minerals. Thus, the relationship between particle size fraction, particle advancing contact angle, and flotation response was explored under well-defined hydrodynamic conditions using model chalcopyrite particles in a 5 dm 3 bottom-driven batch flotation cell. The bottom driven batch flotation cell is commonly used in practice so the methodology can be used in the field with relative ease and may add value to the interpretation of commonly obtained flotation data. The advancing contact angle of individual size fractions in the feed to the flotation stage was measured and controlled to different values. Flotation experiments were carried out under highly reproducible conditions of frother concentration, impeller rotational speed, bubble size, superficial gas velocity, pulp density, ph, and feed particle size distribution. The advancing contact angle was measured on the flotation feed and products using the Washburn method, a method which measures particle wettability by different liquids in a packed column. Thus, the advancing contact angle measured by this method is the surface area weighted average of all particles in the packed column. It was found that the model chalcopyrite particles floated in such a way that suggested their independent behaviour from size fraction to size fraction. Moreover, flotation behaviour was reproducible, within acceptable experimental error, in different tests for the same particle contact angle range and particle size fraction. The flotation behaviour was interpreted as the percentage of a single floating component (called the maximum recovery) after subtraction of recovery by entrainment. This analysis gave a very good fit to the experimental flotation rate data for all particle size fractions and for all contact angle ranges examined. Chalcopyrite particles within the same size fraction and advancing contact angle range were recovered at the same flotation rate with the same floatability component, within acceptable experimental error limits, in different tests. The percentage of the floating component increased with the advancing contact angle in all particle size fractions. The distributed and undistributed rate constants were calculated for each particle size fraction and contact angle range. The distributed rate constant is the flotation rate iv

7 constant of the floatable fraction only, while the undistributed rate constant takes into account both the floatable and non-floatable components. It was observed that the distributed and undistributed rate constants converged at high advancing contact angles (>80 ) and diverged with a decrease in the advancing contact angle due to decreases in the percentage of the floatable component. The undistributed rate constant, and its associated collection efficiency, better described flotation response with respect to particle contact angle and were adopted as the parameters for characterising the floatability of feed particles in the master curves. Evidence for the heterogeneity of the advancing contact angle within each feed size fraction was apparent in the appearance of a significant non-floating component under some conditions. This statement was validated in a parallel study by independent measurements of contact angle distribution in feed size fractions inferred by Time of Flight Secondary Ion Mass Spectrometry (ToF-SIMS) measurements of surface species on individual particles. The correlation between advancing contact angle and ToF-SIMS intensities of hydrophobic and hydrophilic species, allowed individual contact angles on particles to be inferred and the distribution of contact angle values to be determined within a sample population. It was assumed that particles in the non-floating component had advancing contact angles below the critical contact angle required for stable bubbleparticle attachment. This assumption was validated by contact angle measurements on the tailing product which showed that these particles were invariably at or below the critical contact angle under the flotation conditions examined. The flotation results were then used to calculate the critical contact angle as a function of particle size range, assuming a statistical distribution of advancing contact angles about the measured mean value. Generally, and as expected, the critical contact angle for stable bubble-particle attachment increases with particle size above 53 µm, and for particles below 20 m. The recovery and rate constant data were collated into master curves of the undistributed rate constant and collection efficiency as a function of particle size fraction for different contact angle ranges. The contact angle ranges (or bins) were chosen based on the experimental recovery and undistributed rate constant error analysis. Further, the flotation response of chalcopyrite particles conditioned with dicresyl dithiophosphate (DTP) collector was tested, to probe further the effect of different functional groups at the mineral surface. The advancing contact angles of the flotation products were also measured after appropriate sample preparation. The recalculated feed v

8 contact angle values, based on measurements of the flotation products, were in excellent agreement with experimental measurements of the feed itself, a result that confirmed that the particle contact angle did not change significantly through the flotation process. When the flotation response of this different collector system was benchmarked against the calibration curves, the flotation response was the same, within acceptable experimental error, for the same advancing contact angle ranges. Fast floating particles recovered in the first concentrate exhibited higher contact angles than those in subsequent concentrates, while the particles in the final tailings had advancing contact angles at or below the critical value for stable bubble-particle attachment determined on the chalcopyrite-amyl xanthate system. The flotation response with 2% solids (w/w) of chalcopyrite was similar to that at 30% solids (w/w) with chalcopyrite and quartz. Excellent agreement across all size fractions was obtained between the measured feed advancing contact angle on the chalcopyrite-dtp system and the inferred operational contact angle obtained by benchmarking against the calibration curves separately developed on the chalcopyrite-amyl xanthate system. Additionally, the flotation response of copper-activated pyrite, conditioned separately with two different collectors (xanthate and dithiophosphate) was also benchmarked against the calibration curves to probe the effects of mineral specific gravity and functional group under similar physical conditions. Again, the advancing contact angle of the floating component in the concentrates was higher than that of the non-floating component present in the final tailing. The recovery of pyrite, below 80 advancing contact angle measured on the feed, was lower than that of chalcopyrite within the same advancing contact angle range. Thus, the inferred contact angle values of pyrite, based on the benchmark, were lower than the measured feed contact angle values. The critical contact angle values of the pyrite size fractions were determined by fitting a statistical model to the pyrite data alone and were found to be higher than those of the chalcopyrite-amyl xanthate system for particle size fractions greater than 20 m. These observations were attributed to differences in mineral specific gravity and accounted for differences in the floatability of pyrite and chalcopyrite for the same contact angle range and size fraction. Thus, it was concluded that the calibration curves were specific for each mineral specific gravity. Finally, in order to test the validity of the calibration curves, a natural porphyry copper ore with chalcopyrite and pyrite as the principal sulphide minerals was floated with different collector additions of 2, 15, 30, and 40 g/t of the same dithiophosphate collector (DTP). vi

9 These widely different collector additions were chosen to ensure particles achieved discernible differences in hydrophobicity. The contact angle values of the size fractions of the concentrates were measured and the size fractions assayed. The contact angle values of the hydrophobic valuable mineral particles in the concentrate were calculated using the Cassie equation. QEM Scan analysis on the feed size fractions of the ore showed that the bulk of individual copper bearing particles were fully liberated. The gangue components in the concentrate were mainly hydrophobic pyrite and hydrophilic silicates. There was good agreement between the inferred contact angle values of sulphides (chalcopyrite and pyrite combined) based on the chalcopyrite-amyl xanthate system and the calculated feed contact angle values. Advances towards a technique to infer the operational contact angle of specific mineral particles in a natural ore as a function of particle size fraction were progressed as a result of this thesis. vii

10 Table of Contents Acknowledgements... i Publications... ii Abstract... iii Table of Contents... viii List of Figures... xiii List of Tables... xviii 1. Introduction Contact Angle and Floatability Particle Collection Induction Time and Contact Angle Contact Angle and Stability Efficiency Contact Angle Measurement Methods Research Aim, Strategy and Objectives Literature Review Introduction Particle Collection Particle Collection Efficiency Collision Efficiency, E c Attachment Efficiency, E a Stability Efficiency, E s Flotation Kinetics and Floatability Components Wettability and Contact Angle Hydrophobicity and Sulphide Mineral Surface Characteristics Contact Angle and Mineral Surface Roughness Critical Contact Angle and Recovery viii

11 2.6. Contact Angle Measurement Methods Hydrodynamics Power Input and Energy Dissipation Air Flow Rate Superficial Gas Velocity, Bubble Size, and Bubble Surface Area Flux Particle Size Frothers and the Froth Phase in Flotation Summary Experimental Procedures Materials Sample Preparation (Single minerals) Sample Preparation (Natural ore) Methods Contact Angle Manipulation Chalcopyrite-amyl Xanthate System Chalcopyrite-DTP System Pyrite-amyl xanthate/dtp systems Contact Angle Measurement Washburn Technique Equilibrium Capillary Pressure (ECP) Technique Flotation Tests Single Minerals: Chalcopyrite and Pyrite, 2% Solids Chalcopyrite Flotation at 30% Solids Natural Ore Flotation Chemical Assay and Mineralogical Analysis Bubble size distribution measurement Data Analysis Recovery by Entrainment Recovery and Rate Constant Contact Angle and Surface Analysis Surface Analysis: X-ray Photo-electron Spectrometry (XPS) ix

12 Surface Analysis: Time of Flight Secondary Ion Mass Spectrometry (ToF- SIMS) Correlating Particle Size and Contact Angle to the Flotation Response Abstract Introduction Error Analysis Error in Contact Angle Reproducibility of Flotation Tests Contact Angle Variation with Particle Size Contact Angle, Particle Size and Flotation Response Flotation Independence of Mineral Size Fractions The Influence of Gas Flow Rate Contact Angle and Recovery Variation of the Rate Constant with Contact Angle for each Particle Size Fraction Calibration Plots for the Flotation Response Discussion Conclusions Effect of Collector Type, Pulp Density and Particle Specific Gravity on the Flotation Response Abstract Introduction Flotation Response of the Chalcopyrite-DTP System Heterogeneity and Conservation of the Advancing Contact Angle Benchmarking Flotation Response Tests at Higher Solids Percent Flotation Response at High Pulp Density Benchmarking against the Calibration Flotation of Pyrite x

13 Wetting and Dewetting Behaviour of Chalcopyrite and Pyrite Discussion Conclusions Benchmarking a Natural Ore Against Calibration Curves Abstract Introduction Natural Ore Flotation and Inference of an Operational Contact Angle Chalcopyrite Flotation Response Benchmarking Chalcopyrite Flotation Behaviour Sulphide Flotation Response Benchmarking Sulphide Flotation Response Validation of Inferred Operational Contact Angles Surface Analysis Technique for Contact Angle Measurement Advancing Contact Angle Measurements on Heterogeneous Mixtures of Particles Model System Ore System Conclusions Conclusions Future Work References Appendices A1 Material Characterisation A1.1. Chemical assay of chalcopyrite A1.2. Scanning Electron Microscope Images of Chalcopyrite Size Fractions A2 Data Used in Error Analysis And Surface Analysis A2.1. Contact Angle Measurements on Sized And Homogenised Chalcopyrite Samples A2.2. Surface Analysis xi

14 A3 Flotation Response Characterisation A3.1. Flotation Test 1 (Chapter 4) A3.3. Averaged cumulative recovery with time for different contact angle ranges within a size fraction (Chalcopyrite-KAX System)-Chapter A4 Chalcopyrite-KAX and Pyrite-DTP/KAX Model Fitting Procedure (Sections 4.10&5.4.2) A.5.1. KUCC Ore Flotation at 2 g/t Collector Addition A5.2. KUCC Ore Flotation at 15 g/t Collector addition A5.3. KUCC Ore Flotation at 30 g/t Collector Addition xii

15 List of Figures Figure 1.1 Contact angle between bubble and particle in an aqueous medium (Fuerstenau and Raghavan, 1976) Figure 2.1 Schematic representation of a particle of diameter d p sliding around a bubble of diameter d b from a collision or adhesion angle of θ a, to a maximum sliding angle π/2. θ t is the maximum possible collision angle or angle of tangency (Dai et al., 1999) Figure 2.2 Histogram of feed contact angle distribution from ToF-SIMS (a) µm Chalcopyrite sample conditioned with 2g/t DTP, CA=55, (b) µm Chalcopyrite sample conditioned with 2g/t DTP, CA=60. (Brito E Abreu, 2010) Figure 2.3 Particle size as a function of critical surface coverage and contact angle. Methylated quartz particles. (Crawford, 1986) Figure 2.4 Critical contact angle determination. Quartz recovery after 8 minutes of flotation, obtained in a flotation column. (Gontijo et al., 2007) Figure 2.5 Relationship between flotation rate constant k and impeller speed. (Gupta and Yan, 2006) Figure 2.6 Conventional recovery profile with particle size (Pease et al., 2006) Figure 2.7 Comparison of flotation rate constants for ground and spherical ballotini at various sizes (Koh et al., 2009) Figure 2.8 Collision efficiency (E c ), attachment efficiency (E a ) and stability efficiency (E s ) of chalcopyrite particles as a function of particle diameter. Contact angle 71 o, stirring speed 720 rpm, d b = 0.97 mm (Duan et al, 2003) Figure 2.9 Effect of frother concentration on bubble size (schematic) (Cho and Laskowski, 2002) Figure 2.10 Two-phase model description of flotation after (Feteris et al., 1987) Figure 3.1 Experimental outline for single mineral work Figure 3.2 Particle size for 80% passing as a function of grinding time for 2 kg of KUCC ore with 9.3 kg of stainless steel rods (10) Figure 3.3 Wilhelmy Balance for contact angle measurements based on the Washburn technique Figure 3.4 Assembled ECP apparatus ready for contact angle pressure measurement. The apparatus consists of (A) reservoir U-tube, (B) pressure transducer, (C) bleed valve, (D) reducing union and (E) the sample holder. (Stevens, 2005) Figure 3.5 Bottom driven flotation cell with manual scraper for concentrate removal.. 58 Figure 3.6 Cumulative liberation yield as a function of percent liberation for chalcopyrite in KUCC ore feed size fractions Figure 4.1 Material constant, K, of the chalcopyrite particle bed as a function of (a) contact angle for the -20 ( ), ( ), ( ) and +300 (o), µm size fractions, (b) particle size Figure 4.2 Fitted cumulative recovery (%) with time for chalcopyrite size fractions (µm) in flotation tests. (a) Test 1, (b) Test 2, (c) Test ( ), (x), (+), ( ), ( ), ( ), ( ), (o), +300 ( ) xiii

16 Figure 4.3 Maximum recovery as a function of particle size fraction. Size fractions manipulated to different contact angles shown in Table Figure 4.4 Cumulative recovery with time for different contact angle ranges for (a) -20, (b) 20-38, (c) µm chalcopyrite size fractions ( ), ( ), ( ), (+), ( ), (x), ( ) Figure 4.5 Cumulative recovery with time for different contact angle ranges for (a) 53-75, (b) , (c) µm chalcopyrite size fractions ( ), ( ), ( ), (+), ( ), (x), ( ) Figure 4.6 Cumulative recovery with time for different contact angle ranges for (a) , (b) , (c) 300+ µm chalcopyrite size fractions ( ), ( ), ( ), (+), ( ), (x), ( ) Figure 4.7 Variation of the distributed ( ) k, and undistributed (O) k*, rate constants with contact angle and particle size fraction. J g =0.3 cm/s, d b =0.48 mm, 1200 rpm, 2% solids. Fitted lines are calculated using equation and parameters shown in Table Figure 4.8 Maximum recovery as a function of particle size for different contact angle ranges. Chalcopyrite single mineral. Agitation Speed=1200 rpm, Jg=0.3 cm/s Figure 4.9 Undistributed rate constant as a function of particle size for different contact angle ranges. Chalcopyrite single mineral. Agitation speed, 1200 rpm, Jg=0.3 cm/s Figure 4.10 Collection efficiency (E coll ) as a function of particle size for different contact angle ranges. Chalcopyrite single mineral. Agitation speed, 1200 rpm, Jg=0.3 cm/s Figure 4.11 Contact angle distribution in a size fraction of the feed ( µm, Z crit =55 ). Gamma distribution, measured contact angle = mean (Mean contact angle value indicated on the top left hand corner). The area to the left of Z crit (vertical broken line) represents the non-floating component in each contact angle range. The amount of the nonfloating fraction decreases, and R max increases, as the mean contact angle increases Figure 4.12 Experimental maximum recovery versus Gamma distribution model maximum recovery. R 2 = Figure 4.13 Experimental versus Model unsized maximum recovery at feed contact angles (mean) of 38 ( ), 53 ( ), 63 ( ), 68 ( ), 73 ( ), 78 (O), 88 ( ), R 2 = Figure 5.1 Cumulative recovery with time for chalcopyrite size fractions conditioned with DTP and floated at 2% solids (a) Test 1 (b) Test 2. The continuous lines are for equation 3.10, after subtraction of recovery by entrainment Figure 5.2 Inferred operational advancing contact angle compared with the measured feed advancing contact angle of DTP conditioned chalcopyrite size fractions with (a) low, R 2 =0.94 (Test 1) (b) high advancing contact angle (Test 2) regimes, R 2 = Figure 5.3 Advancing contact angles of concentrates and tails for DTP conditioned chalcopyrite with (a) low (2 g/t DTP, Test 1), (b) high (20 g/t DTP, Test 2) advancing contact angles compared with the critical advancing contact angle curve (chalcopyriteamyl xanthate system) Figure 5.4 Cumulative recovery with time for chalcopyrite size fractions conditioned with DTP, floated at 30% solids (w/w), (a) Test 3 (b) Test 4. The continuous lines are for equation 3.10, after subtraction of recovery by entrainment xiv

17 Figure 5.5 Operational advancing contact angle compared with measured feed advancing contact angle for chalcopyrite feed size fractions floated at 30% solids (w/w), (a) Test 3, (b) Test 4. R 2 =0.98 in both tests Figure 5.6 Cumulative recovery with time for pyrite size fractions (a) KAX, Test 5 - low contact angle regime, (b) KAX, Test 6 - high contact angle regime, (c) DTP, Test 7. The continuous lines are for equation 3.10, after subtraction of recovery by entrainment Figure 5.7 Comparison of maximum recoveries (%) for chalcopyrite and pyrite for each size fraction. The values are for advancing feed contact angles for pyrite and chalcopyrite in Table Figure 5.8 Inferred operational advancing contact angle compared with measured feed advancing contact angle for (a) pyrite-xanthate system with low advancing contact angle Test 5, (b) pyrite-xanthate system with high advancing contact angle Test 6, (c) pyrite- DTP system with low advancing contact angle Test Figure 5.9 Variation of the mean and mode advancing contact angle values for a sample, assuming gamma statistical distribution of individual particle contact angles. The mean is the value measured by the Washburn technique. The mean and mode contact angle values determine the contact angle distribution in a feed sample Figure 5.10 Experimental versus model maximum recovery for pyrite flotation, 2% solids. (a) Assumption 1, critical advancing contact angle for pyrite is the same as for chalcopyrite, R 2 =0.84, (b) Assumption 2, pyrite has a different critical advancing contact angle than chalcopyrite, R 2 = Figure 5.11 Cumulative mass recovered into the tail, second, and third concentrate (total = recalculated feed) against the cumulative mean contact angle for pyrite (O) and chalcopyrite ( ) for (a) -20 (Tests 2 and 6), (b) (Tests 2 and 6), (c) (Tests 1 and 6), (d) +300 (Tests 2 and 6) µm size fractions from Tables 5.1 and 5.3. Tests compared with similar mean contact angles values in the feed Figure 5.12 Variation of the critical advancing contact angle with particle size for pyrite and chalcopyrite Figure 6.1 Maximum recovery (R max ) of chalcopyrite as a function of particle size at different collector additions for KUCC ore. Collector (DTP) added at 2 ( ), 15 ( ), 30 ( ), and 40 g/t (o). Frother 37.5 g/t MIBC, 1200 rpm, 30% solids (w/w) Figure 6.2 Cumulative Cu recovery with time, natural copper ore, (a) 2 g/t, (b) 15 g/t, (c) 30 g/t dicresyl dithiophosphate collector. For size fractions:( ) -20, ( ) 20-38, ( ) 38-53, (+) 53-75,( ) , ( ) , ( ) , (o) , ( ) +300 The continuous lines are for equation Recovery by entrainment has been subtracted Figure 6.3 Collection efficiencies of particle size fractions for 2, 15, 30 g/t ore flotation tests benchmarked against calibration curves of the collection efficiency, E coll, against particle size for different contact angle ranges (from Chapter 4, Figure 4.10) Figure 6.4 Maximum recovery (R max ) of sulphide minerals (chalcopyrite and pyrite) as a function of particle size at different collector additions 2 ( ), 15 ( ), and 30 ( ) g/t DTP for KUCC ore Figure 6.5 Cumulative sulphur recovery with time, natural copper ore, (a) 2 g/t, (b) 15 g/t, (c) 30 g/t dicresyl dithiophosphate collector. For size fractions:( ) -20, ( ) 20-38, ( ) xv

18 38-53, (+) 53-75,( ) , ( ) , ( ) , (o) , ( ) +300 The continuous lines are for equation Recovery by entrainment has been subtracted Figure 6.6 A comparison of the inferred operational contact angle of chalcopyrite and sulphide (chalcopyrite and pyrite) for different size fractions at collector additions of (a) 2 g/t, (b) 15 g/t, (c) 30 g/t DTP Figure 6.7 Distribution of the advancing contact angles measured using ToF-SIMS of chalcopyrite particles in the concentrate of the µm size fraction of the test at 15 g/t Figure 6.8 Advancing contact angle as a function of the volume fraction of hydrophobic chalcopyrite in a mixture of clean quartz (zero contact angle) and chalcopyrite µm size fraction for both minerals Figure 6.9 Liberation of copper sulphides (chalcopyrite) in selected coarse particle size fractions of the concentrate produced with15 g/t DTP Figure 6.10 Particle maps for the ore (a) feed, (b) concentrate, and (c) tailing, for the µm size fraction determined by QEM-Scan. Flotation test at 15 g/t collector addition. NSG is non-sulphide gangue excluding silicates. Cu Sulphides is chalcopyrite. Fe Sulphides is pyrite Figure 6.11 Comparison of the inferred operational and measured contact angle of sulphide minerals in a natural ore. R 2 = Figure A1.1 SEM microgram of -20 µm chalcopyrite size fraction Figure A1.2 SEM microgram of µm chalcopyrite size fraction Figure A1.3 SEM microgram of µm chalcopyrite size fraction Figure A1.4 SEM microgram of µm chalcopyrite size fraction Figure A1.5 SEM microgram of µm chalcopyrite size fraction Figure A1.6 SEM microgram of µm chalcopyrite size fraction Figure A1.7 SEM microgram of µm chalcopyrite size fraction Figure A1.8 SEM microgram of µm chalcopyrite size fraction Figure A1.9 SEM microgram of +300 µm chalcopyrite size fraction Figure A2.1 XPS spectra for -20 µm size fraction, surface atomic concentrations Figure A2.2 XPS spectra for -20 µm size fraction, Cu species Figure A2.3 XPS spectra for -20 µm size fraction, Fe species Figure A2.4 XPS spectra for -20 µm size fraction, O species Figure A2.5 XPS spectra for -20 µm size fraction, C species Figure A2.6 XPS spectra for -20 µm size fraction, S species Figure A2.7 XPS spectra for µm size fraction, surface atomic concentration Figure A2.8 XPS spectra for µm size fraction, Cu species Figure A2.9 XPS spectra for µm size fraction, Fe species Figure A2.10 XPS spectra for µm size fraction, O species Figure A2.11 XPS spectra for µm size fraction, C species Figure A2.12 XPS spectra for µm size fraction, S species Figure A2.13 XPS spectra for µm size fraction, surface atomic concentration Figure A2.14 XPS spectra for µm size fraction, Cu species Figure A2.15 XPS spectra for µm size fraction, Fe species xvi

19 Figure A2.16 XPS spectra for µm size fraction, O species Figure A2.17 XPS spectra for µm size fraction, C species Figure A2.18 XPS spectra for µm size fraction, S species Figure A2.19 XPS spectra for µm size fraction, surface atomic concentration Figure A2.20 XPS spectra for µm size fraction, Cu species Figure A2.21 XPS spectra for µm size fraction, Fe species Figure A2.22 XPS spectra for µm size fraction, O species Figure A2.23 XPS spectra for µm size fraction, C species Figure A2.24 XPS spectra for µm size fraction, S species Figure A2.25 XPS spectra for µm size fraction, surface atomic concentration. 182 Figure A2.26 XPS spectra for µm size fraction, Cu species Figure A2.27 XPS spectra for µm size fraction, Fe species Figure A2.28 XPS spectra for µm size fraction, O species Figure A2.29 XPS spectra for µm size fraction, C species Figure A2.30 XPS spectra for µm size fraction, S species Figure A2.31 XPS spectra for µm size fraction, surface atomic concentration Figure A2.32 XPS spectra for µm size fraction, Cu species Figure A2.33 XPS spectra for µm size fraction, Fe species Figure A2.34 XPS spectra for µm size fraction, O species Figure A2.35 XPS spectra for µm size fraction, C species Figure A2.36 XPS spectra for µm size fraction, S species Figure A2.37 XPS spectra for µm size fraction, surface atomic concentration Figure A2.38 XPS spectra for µm size fraction, Cu species Figure A2.39 XPS spectra for µm size fraction, Fe species Figure A2.40 XPS spectra for µm size fraction, O species Figure A2.41 XPS spectra for µm size fraction, C species Figure A2.42 XPS spectra for µm size fraction, S species Figure A2.43 XPS spectra for µm size fraction, surface atomic concentration Figure A2.44 XPS spectra for µm size fraction, Cu species Figure A2.45 XPS spectra for µm size fraction, Fe species Figure A2.46 XPS spectra for µm size fraction, O species Figure A2.47 XPS spectra for µm size fraction, C species Figure A2.48 XPS spectra for µm size fraction, S species Figure A2.49 XPS spectra for +300 µm size fraction, surface atomic concentration Figure A2.50 XPS spectra for +300 µm size fraction, Cu species Figure A2.51 XPS spectra for +300 µm size fraction, Fe species Figure A2.52 XPS spectra for +300 µm size fraction, O species Figure A2.53 XPS spectra for +300 µm size fraction, C species Figure A2.54 XPS spectra for +300 µm size fraction, S species xvii

20 List of Tables Table 2.1 Calculated values of contact angles (θ) of n-alkyl groups (Somasundaran and Wang, 2006) Table 3.1 Chemical assays (%) of pyrite size fractions Table 3.2 Chemical Assays (wt %) of chalcopyrite size fractions*used in the tests Table 3.3 Size distribution for 2 kg of KUCC ore ground for 12 minutes (10 rods, 9.3 kg) Table 3.4 Contact angle manipulation for chalcopyrite size fractions Table 3.5 Contact angle manipulation for copper activated pyrite using KAX and DTP Table 3.6 Size distribution of flotation feed (chalcopyrite and pyrite) Table 3.7 Bulk mineralogical assay (Weight %) for KUCC ore by QEM Scan Table 3.8 Sample calculation of recovery by entrainment and by true flotation for the -20 µm size fraction Table 3.9 Experimental and fitted recoveries with time for +300 µm size fraction with 90 contact angle Table 4.1 Error in advancing contact angle measurement on chalcopyrite treated with potassium amyl xanthate (200 and 50 g/t, respectively for the high and low contact angle regimes, all size fractions conditioned together; Section ), n= Table 4.2 Volume of measuring liquid imbibed by 2 g of chalcopyrite conditioned with potassium amyl xanthate (200 and 50 g/t, respectively for the high and low contact angle regimes, all size fractions conditioned together; Section ) Table 4.3 Error in R max and k* for the same chalcopyrite size fractions with high and low contact angles as in Tables 4.1 and 4.2, n=3. All size fractions conditioned and floated together Table 4.4 Contact angle and XPS atomic concentrations (%) (Appendix A2.2) on surfaces of chalcopyrite (in the absence of collector) size fractions. All particles conditioned together and then separated for contact angle and XPS measurement Table 4.5 Regression coefficients for contact angle and XPS atomic concentrations Table 4.6 Feed advancing contact angles and R max of manipulated chalcopyrite size fractions, Jg=0.3 cm/s Table 4.7 Advancing contact angles ( ) and undistributed rate constant k* (1/min) of chalcopyrite size fractions in different tests, Jg=0.3 cm/s Table 4.8 Contact angle, flotation recovery and rate constants for selected size fractions. Effect of changing J g Table 4.9 Empirical equations for the calculation of contact angle from the undistributed rate constant, by size fraction. Contact Angle (±2.5 o ) = a(k*) b Table 4.10 Empirical Equations relating the contact angle and the collection efficiency, E coll. Contact angle (±2.5 o ) = a(e coll ) b Table 4.11 Calculated critical contact angle, Z crit, using the Gamma distribution model for contact angle values around the nominal (measured) value xviii

21 Table 5.1 Contact angle values of flotation products and the measured and recalculated advancing contact angles ( ) for DTP conditioned chalcopyrite before and after flotation, and the corresponding maximum recovery and undistributed rate constants. Low (Test 1) and high (Test 2) advancing contact angle regimes. (Table 3.4; Section ) Table 5.2 Flotation response of chalcopyrite with quartz at 30% solids (w/w). Chalcopyrite conditioned with DTP (2 and 20 g/t) prior to flotation. Size fractions conditioned individually, and floated together. (Section ; Table 3.4) Table 5.3 Advancing contact angles (CA) of pyrite flotation feed and products (First, second concentrates, and tail), and corresponding maximum recovery and undistributed rate constants. Corresponding R max values for chalcopyrite size fractions within the same advancing contact angle range as in Tests 5-7. (Section ; Table 3.5) Table 5.4 Advancing and receding contact angles of feed chalcopyrite and pyrite conditioned with KAX, µm size fractions. (Section ) Table 5.5 Pyrite flotation data model fitting to a gamma statistical distribution with mean (measured) and mode contact angle. The critical contact angle, fixed for each size fraction, was allowed to vary to obtain the best fit Table 6.1 Collection efficiency, E coll, and inferred operational contact angles of chalcopyrite in KUCC ore feed after different collector additions Table 6.2 Cu and S Assays of KUCC ore concentrates from flotation tests with calculated weight % and volume fractions of chalcopyrite (chp), pyrite (Pyke et al.) and non-sulphide gangue (NSG) Table 6.3 Collection efficiency, E coll, and inferred operational contact angles (CA) of sulphides (chalcopyrite and pyrite) in KUCC ore feed after different collector additions Table 6.4 Contact angles measured on the concentrates, the calculated sulphide contact angle values within concentrate, and the back calculated feed contact angles. The back calculated feed contact angles use the R max values. ND not determined Table A1.1 Chalcopyrite assay, wt%, before (B) and after (A) a pre-flotation cleaning flotation stage Table A2.1 Repeated Contact angle measured on homogenised chalcopyrite sample (conditioned with 200 g/t potassium amyl xanthate), high contact angle Table A2.2 Maximum recovery, (%) and undistributed rate constant (min -1 ) for homogenised sample with contact angle in Table A Table A2.3 Repeated Contact angle measured on homogenised chalcopyrite sample (conditioned with 50 g/t potassium amyl xanthate), Low contact angle Table A2.4 Maximum recovery, (%) and undistributed rate constant (min -1 ) for homogenised sample with contact angle in Table A Table A2.5 Advancing contact angle of collectorless chalcopyrite size fractions Table A3.1 Feed size distribution and contact angle Table A3.2 Mass recovery (g) of concentrates, tail, and water Table A3.3 Size distribution of flotation products and the recalculated feed Table A3.4 Recovery by size, % Table A3.5 Recovery (%), after subtraction of entrainment xix

22 Table A3.6 Cumulative recovery ( %) with flotation time Table A3.7 Cumulative recovery with time for flotation Tests 2-7 (Chapter 4) Table A3.8 Averaged cumulative recovery with time from different tests for different contact angle ranges within a size fraction, µm size fractions. Data used in the generation of calibration curves Table A3.9 Averaged cumulative recovery with time from different tests for different contact angle ranges within a size fraction, µm size fractions. Data used in the generation of calibration curves Table A3.10 Cumulative recovery with time for flotation Tests 1-7 (Chapter 5) Table A4.1 Gamma statistical distribution model fit to chalcopyrite contact anglerecovery data, -20 µm size fraction Table A4.2 Gamma statistical distribution model fit to chalcopyrite contact anglerecovery data, µm size fraction Table A4.3 Gamma statistical distribution model fit to chalcopyrite contact anglerecovery data, µm size fraction Table A4.5 Gamma statistical distribution model fit to chalcopyrite contact anglerecovery data, µm size fraction Table A4.6 Gamma statistical distribution model fit to chalcopyrite contact anglerecovery data, µm size fraction Table A4.7 Gamma statistical distribution model fit to chalcopyrite contact anglerecovery data, µm size fraction Table A4.8 Gamma statistical distribution model fit to chalcopyrite contact anglerecovery data, µm size fraction Table A4.9 Gamma statistical model fit to chalcopyrite contact angle-recovery data, +300 µm size fraction Table A4.9 Gamma statistical model fit to pyrite contact angle-recovery data Table A5.1 Head ore assays Table A5.2 Assays of concentrates and tail, wt% Table A5.3 Cu recovery calculation Table A5.4 Sulphide recovery calculation Table A5.5 Assays of concentrate and tail, wt % Table A5.6 Mass recovery and recalculated feed distribution and assays Table A5.7 Cu recovery calculation Table A5.8 Sulphur Recovery, wt% Table A5.9 Concentrate and tail assays Table A5.10 Mass Recovery and recalculated feed size distribution, Cu, and S assays Table A5.11 Cu Recovery Calculation Table A5.12 Sulphur Recovery Calculation xx

23 CHAPTER ONE 1. Introduction Flotation is the process of separating one mineral type from another by selective gas bubble attachment onto one of the mineral surfaces, while the particles are submerged in a liquid medium (Novich, 1990). Flotation separation relies on differences in surface properties (Prestidge and Ralston, 1996b), either naturally occurring or induced by addition of surfactants. Particles are recovered via two primary mechanisms. The first mechanism is by true flotation, in which hydrophobic particles attach to air bubbles and the bubbleparticle aggregates rise to the froth phase, and are collected to the concentrate launder via the froth phase. The second mechanism is by entrainment, which involves the nonselective recovery of particles in the water recovered to the concentrate irrespective of whether they are hydrophobic or hydrophilic (Harris et al., 2002). Recovery by entrainment is subtracted from the total recovery to obtain recovery by bubble-particle attachment (Savassi et al., 1998) Contact Angle and Floatability Mineral surfaces can be characterised as being hydrophobic (water repelling), hydrophilic (water loving ), or between these two extremes. The degree of hydrophobicity and wetting characteristics may be represented by the static contact angle in the first instance (Nadkarni and Garoff, 1992). Wetting refers to the way a liquid on a solid substrate spreads out or wets the surface. The static contact angle is a thermodynamic quantity defined geometrically as the angle between planes tangent to the surfaces of the solid and the liquid at the wetting perimeter where liquid, solid, and gas phases intersect (Fuerstenau and Raghavan, 1976) (Figure 1.1). 1

24 Figure 1.1 Contact angle between bubble and particle in an aqueous medium (Fuerstenau and Raghavan, 1976). The contact angle forms the basic thermodynamic criterion for separability of minerals by flotation. Flotation recovery increases with the contact angle of particles in the pulp (Prestidge and Ralston, 1996a), reaching a peak at the maximum contact angle of the collector-mineral system (Novich, 1990). The recovery of hydrophobic particles by attachment to air bubbles is widely regarded as a first order kinetic process with respect to particle concentration in the cell. For first order kinetics to apply, there should be enough bubble surface, i.e., the bubble surface is not overloaded with particles, and the number of bubbles remains constant. This ensures the bubble surface is not rate limiting or rate determining. In addition, the particles must float independently of each other, with minimum or no interaction so that individual particle recovery rate is not influenced by other particles in the pulp (Imaizumi and Inoue, 1963; Tomlinson and Fleming, 1963b; Harris and Chakravarti, 1970; Harris and Cuadros-Paz, 1978; Harris et al., 2002). Past investigations on interaction of particles in the pulp used either a mixture of different minerals (one size fraction each) with different sizes (Harris and Cuadros-Paz, 1978), or a narrow range of particle size in a low solids content pulp (Imaizumi and Inoue, 1963). Further, the degree of hydrophobicity of the particles used did not cover the whole range commonly encountered in practical flotation systems. In this thesis, the possible solid-solid interaction of different particle size fractions with each other when the contact angle is varied is investigated. If indeed, particles float independently of each other when conditions for first order kinetics are met, then it should be possible to study the effect of contact angle on flotation for a polydispersed particulate system when the hydrodynamic conditions are kept constant. This possibility is explored further below. 2

25 The study of the flotation response of mineral particles has been characterised by an empirical floatability parameter, with floatability being defined as the propensity of a mineral particle in a flotation pulp to be collected by true flotation, and reflects the interaction of the mineral properties with the pulp chemical environment, (Harris et al., 2002). Floatability is affected by particle size, density, shape, mineralogical composition, the degree of surface oxidation, type and extent of reagent coverage, and degree of aggregation of the particles in the system (Harris et al., 2002). The foregoing is an elaborate list of the factors affecting the floatability of particles. It is obvious that floatability as a parameter is difficult to quantify, let alone assign a value to each of those listed factors. And, as noted by the authors, most flotation streams cannot be described by a single floatability value due to the large number of variables involved, but an array of floatabilities (Imaizumi and Inoue, 1963; Harris and Chakravarti, 1970; Harris et al., 2002). With these observations and apparent difficulties of using floatability parameters to characterise flotation response, this work seeks to use relatively easy to measure and quantifiable parameters such as the particle contact angle and particle size to characterize flotation behaviour. The motivation for this project is that when valuable mineral particles are sized into reasonably narrow size ranges (e.g., 2 series of sieves), the particle size, density, shape, and mineralogical composition can be assumed to be the same for each size fraction for a given mineral. The floatability of mineral particles has also been shown to be subject to a critical contact angle (Crawford and Ralston, 1988; Gontijo et al., 2007). All particles of a given size fraction with contact angles above the critical value for that size may be recovered while those particles with lower contact angles may not form a stable bubble-particle aggregate. Combining the prospects of a statistical distribution of contact angles within a given feed size fraction with the concept of the critical contact angle, it may be possible to predict the maximum recovery that can be realised for any given size fraction of known contact angle in the feed under well-defined hydrodynamic conditions. This possibility is explored in this thesis. 3

26 1.2. Particle Collection The key steps in the collection of particles by bubbles in the pulp involve bubble-particle collision, attachment and detachment (Dai et al., 1998; Bloom and Heindel, 2003; Duan et al., 2003). The collection of particles is represented by a collection efficiency, E coll, which is a product of the three sub-processes: (1.1) where E c, E a, and E s are the collision, attachment and stability efficiencies, respectively, between bubbles and particles in the pulp (Derjaguin and Dukhin, 1961; Ralston et al., 1999b). Hydrodynamic interactions (bubble/particle sizes, bubble velocity, turbulence) are dominant in the collision efficiency while interfacial forces (e.g., particle hydrophobicity) play a major role in the attachment efficiency. The stability of the bubble-particle aggregate (a function of both hydrodynamics and interfacial forces) against disruptive forces in the pulp determines the stability efficiency (Duan et al., 2003). A froth recovery factor, E f, may be incorporated into the above equation if the froth height is significant (Harris et al., 2002). In laboratory flotation cells, the froth effect is assumed negligible when the flotation process is carried out at shallow froth depths (Vera et al., 1999). Each of the above sub-processes has a bearing on the flotation rate constant. A model that brings together both the hydrodynamic and interfacial factors has been proposed, bringing together the parameters critical to the recovery of minerals. The flotation rate constant, k, is given by (Duan et al., 2003; Pyke et al., 2003) G fr 0.33 db b 1 k Ec. Ea. 1 db. Vcell fl v 3 b E s (1.2) mechanical term primary turbulence term elementary subprocesses where G fr is the gas flow rate, d b is the bubble diameter, V cell is the volume of the flotation cell, is the turbulent mean energy dissipation, is the difference between particle and fluid densities, is the fluid density, is fluid viscosity, fl b is the bubble velocity. When the hydrodynamic conditions are relatively constant, the flotation rate depends on efficiency of the last three sub-processes, whose product gives the collection efficiency, 4

27 E coll. Under steady state conditions, the collision efficiency may be assumed constant, so that any changes to the rate constant can be attributed to changes in the attachment and stability efficiencies. It is worthwhile, however, to mention in passing, the impact of some parameters on the collision efficiency. The collision efficiency between bubbles and particles increases as particle size, density and energy input increase, due to the rapid increase in particle momentum resulting in a higher level of bubble-particle collisions (Duan et al., 2003; Koh and Schwarz, 2003). A reduction in bubble size increases the number of bubbles in the pulp, thereby increasing the collision frequency Induction Time and Contact Angle When a particle collides with a bubble in the pulp zone, the liquid separating the two must be displaced if attachment is to occur. The induction time is the minimum time necessary for the attachment of an air bubble to the mineral surface (Gu et al., 2003; Gu et al., 2004). The water film between a mineral surface and air bubble must thin to a critical value, rupture, and recede (drainage) followed by formation of a three-phase wetting perimeter before successful attachment can take place (Ralston et al., 1999a). The rate of movement of the three-phase contact (TPC) line depends on the surface chemistry and roughness of the particle (Krasowska et al., 2006). The surfaces of mineral particles are known to be heterogeneous in terms of chemistry (Piantadosi et al., 2000), and roughness (Sedev et al., 2004; Krasowska et al., 2009). These heterogeneities may result in differences in wetting behaviour (Miller et al., 1996). Surface roughness and high hydrophobicity shorten the induction time and attachment may take place in the first collision (Krasowska and Malysa, 2007; Najafi et al., 2008). Surface roughness may assist in the thinning and rupturing of the intervening liquid film, shortening the attachment time and increasing the attachment efficiency. This result may have implications on the flotation rate constants of mineral particles with different shapes and roughness Contact Angle and Stability Efficiency The stability efficiency, as mentioned before, is dependent on both the hydrodynamics and interfacial forces (the contact angle). The strength of a bubble-particle bond (attachment strength) is dependent on the contact angle. The larger the contact angle, the stronger is the bubble-particle aggregate bond (Laskowski, 1986), and the larger the contact angle, the larger the detachment force (Schimann, 2004), and hence the more stable the aggregate is against turbulent disruptive forces in the cell. When the cell hydrodynamic conditions are 5

28 constant, the attachment strength, particle size and specific gravity may cause variations in the recovery and rate constants. The flotation rate decreases with an increase in specific gravity of mineral (Polat and Chander, 2000), most likely due to increased mass of particles for the same size, as it affects collection efficiency. In the case of coal flotation, the flotation rate first increased and then decreased as the particle size increased, indicating an optimum size for flotation. With an increase in specific gravity of the mineral, the particle size for maximum rate of flotation decreased to lower values (Polat and Chander, 2000). Therefore, the increase in the mass of particles either as a result of an increase in size or specific gravity results in changes in the flotation response Contact Angle Measurement Methods Although the contact angle of mineral particles forms the basis of separation from gangue in the pulp, there is currently no method to determine the value under real flotation conditions. Methods available to measure contact angles of particles may not be suitable in a situation where the mineral of interest is low in concentration in a gangue matrix. The measurement of contact angles on powder materials uses the capillary method, in which the penetration of the measuring liquid into a compacted powder bed is characteristic of the wettability of the powder material. The most common methods are the Washburn (Washburn, 1921) and Equilibrium Capillary Pressure (ECP) (Diggins and Ralston, 1993; Stevens, 2005) techniques. The ECP technique was developed to directly measure the advancing contact angle of mineral particles (Diggins and Ralston, 1993) and later adapted to measure the receding contact angle (Stevens, 2005). The Washburn and Equilibrium Capillary Pressure techniques have been used to measure contact angles on homogeneous and heterogeneous particle beds with some success (Stevens, 2005; Priest et al., 2008). However, the concept and use of the contact angle as a thermodynamic parameter has remained largely academic, with no direct application in a process stream. Thus, the extent of the influence of the contact angle on flotation recovery and rate constant is not fully understood or appreciated in practice (Chau et al., 2009). This state of affairs has developed from the realisation that natural ores have low proportion of valuable minerals that are mixed with hydrophilic gangue so that direct contact angle measurements on such a mixture would not yield a meaningful result. On the other hand, ore concentrates have a significantly higher proportion of target mineral so that contact angle measurements on these particles should give some clue as to the degree of 6

29 hydrophobicity achieved by the particles when conditioned with collector. The difficulty is in the determination of the actual contact angles of the valuable mineral, given that some gangue minerals are also recovered, either as liberated particles or as composites. It is anticipated that concentrates from natural ore flotation may be good candidates from which to extract contact angle data that may be linked to the rate at which particles of a given size fraction are recovered by bubble-particle attachment. This thesis wrestles with the challenge of inferring the contact angle of hydrophobised feed particles based on the rate of flotation under well-defined hydrodynamic conditions. The present study is an attempt to develop a method that can be used to infer the contact angle of particles of sizes normally encountered in flotation. The inferred contact angle is an operational contact angle that can be used to compare the response of mineral particles in an ore to collector conditioning. The method is based on flotation response, a potentially true indicator of the degree of hydrophobicity of particles in the pulp. The approach involves the use of preconditioned feed particles (Harris and Cuadros-Paz, 1978) of known contact angles coupled with short conditioning times before flotation to limit the possibility of surface changes prior to flotation, and direct measurement of contact angle of the flotation products to demonstrate conservation of the contact angle through flotation and to demonstrate particle contact angle heterogeneity Research Aim, Strategy and Objectives The overall aim of this project is to develop a method to infer an operational contact angle of particles in the pulp. In this thesis, the flotation response (in terms of maximum recovery and rate constant) of chalcopyrite is characterised as a function of particle size and contact angle. The adopted approach is to extend the work of characterizing floatability through the measurement of contact angle on model particles (single mineral chalcopyrite) in the feed to a flotation cell as a means to effectively calibrate the flotation separation itself, both as a function of particle contact angle and size. This methodology and its application are further outlined below. Flotation is carried out under well-characterised conditions of superficial gas rate, pulp density, bubble size, and agitation speed in a laboratory flotation cell. A series of rate constant versus particle size (with varying contact angles) calibration curves are obtained under the same experimental conditions. This stage is referred to as calibration because it 7

30 uses high purity, well-characterised, fully liberated chalcopyrite. Initially, these calibration curves are obtained for chalcopyrite single minerals floated without gangue. This is followed by flotation with gangue (quartz) to establish the effect of solids content on the flotation response, as most practical feeds always have a significant solids content. Subsequent flotation tests are carried out with a different collector and a different mineral (e.g., pyrite). The different mineral-collector systems serve to validate observations on chalcopyrite and invoke confidence in the methodology. A validation step is included in this work and involves the flotation of a natural copper ore from Kennecott Utah Copperton Concentrator (KUCC), USA. The ore is floated under the same hydrodynamic conditions as the single mineral work and at different collector additions, and the concentrates are sized, dried, and contact angles measured. It is proposed to develop an ore floatability characterization tool (calibrated flotation cell) that is calibrated using parameters of particle size and contact angle. In turn, this calibrated flotation cell may be used to infer an operational contact angle of hydrophobic particles as a function of particle size in an ore. The research strategy and stages of this project can be stated thus: Characterize the flotation response of relevant single minerals in terms of contact angle and particle size. It is anticipated that the contact angle of particles of different size fractions may be different even when subjected to the same treatment due to issues of surface oxidation and roughness, which may vary with particle size, Generate calibration curves of rate constant and collection efficiency against particle size for specific contact angle ranges under well-defined hydrodynamic conditions, Test the validity of the calibration at a pulp density similar to that used in plant practice, and also benchmark the flotation response of a different mineral-collector system (pyrite) and another chalcopyrite-collector system, Benchmark (compare flotation response) an ore against the calibration to infer an operational particle contact angle by size, 8

31 Independently confirm the inferred particle contact angle by direct measurements on flotation feed and products, and indirectly by Time of Flight Secondary Ion Mass Spectrometry (ToF-SIMS). The objectives of this thesis are as follows: To demonstrate reliability of contact angle measurements on sized particle beds. This is necessary to establish the magnitude of error in the Washburn technique across all size fractions, Characterize the flotation behaviour of mineral size fractions with controlled contact angle so as to generate curves of the rate constant and collection efficiency against particle size for different contact angle ranges, Model the maximum recovery (i.e., floatability component) based on contact angle and size in order to calculate the critical contact angle for stable bubble-particle attachment under constant hydrodynamic conditions, Demonstrate the conservation of floatabilities within a size fraction based on contact angle measurements before and after flotation, as a way to show that the contact angle remains unchanged (within experimental error) in flotation and sample preparation, Demonstrate that there is a heterogeneity of contact angle values which exist in the feed and which gives rise to particles reporting to the flotation products at different rates. This is a key point as it demonstrates the basis of the flotation process, Demonstrate the possibility of inferring an operational contact angle from the flotation response for a natural ore. 9

32 CHAPTER TWO 2. Literature Review 2.1. Introduction This thesis develops a method that can be used to infer an operational contact angle of particles of a given size fraction in the pulp based on the rate of recovery in flotation. In this chapter, the important aspects of the collection of particles by bubbles in a flotation cell are reviewed from the current body of knowledge. The collision, attachment, and detachment processes between bubbles and particles, and their impact on the flotation rate and recovery are reported. The effects of particle surface chemistry and roughness on the contact angle and bubble-particle attachment are discussed. The role of the contact angle, how it is measured, and its correlation to flotation rate and recovery is elucidated, showing the existence of a critical contact angle for stable bubble-particle attachment. This is followed by an analysis of the role of gas dispersion parameters, hydrodynamics and froth effects on the flotation response of particles. Finally, the possible approaches, assumptions, and hypotheses to be addressed are discussed Particle Collection The flotation process involves the collection of hydrophobic particles by bubbles from an agitated pulp. The conditions for optimum operation vary from plant to plant depending on ore type and the reagents in use. Common in flotation science is the selective hydrophobisation of one mineral surface (usually the valuable mineral) while keeping the other (gangue) hydrophilic. Air bubbles preferentially attach to hydrophobic mineral surfaces forming a bubble-particle aggregate of lower density than the mineral particle on its own. The aggregate rises to the froth phase where the particles are collected as a concentrate. An understanding of the stages in the particle collection process is important if successful optimisation is to be realised. 10

33 The rate of particle recovery from the pulp is proportional to the number of floatable particles remaining and is given by (Pyke et al., 2003) dn dt p kn p Z pb E coll (2.1) where t is flotation time, k is the flotation rate constant, N p is the number of floatable particles, E coll is the collection efficiency, and Z pb is the collision frequency per unit volume between particles and bubbles of diameters d p and d b respectively, and is given by Z pb N p Nbd pb V p V b (2.2) where d pb = (d p +d b /2), and, is the root mean square velocity fluctuations of particles and bubbles and is given by (2.3) where is the mean energy dissipation, is the difference between particle and fluid densities, fl is the fluid density, is fluid viscosity. This equation was derived for particles and validated by Brady et al., (2006), using particle image velocimetry. Bloom and Heindel, (2002) used the same expression to calculate velocities for both particles and bubbles. When the bubble diameter is much greater than the particle diameter, the particle diameter and its velocity may be neglected and the collision efficiency may be approximated by Z pb 5N N p b db d 7 9 b fl 2 3 (2.4) The flotation rate constant then becomes k 5N b db d 7 9 b fl 2 3 E coll (2.5) 11

34 N b can be calculated from the gas flow rate, G fr, and the residence time, t r of bubbles in the unit volume: N b 6G 3 b fr d. V cell t r (2.6) where V cell is the volume of the flotation cell. The residence time can also be expressed as the time that the bubbles of velocity v b remain in the unit volume (Duan et al., 2003): t r = (1 unit length)/v b The final form of the flotation model is given by (Duan et al., 2003): G fr 0.33 db b 1 k Ec. Ea. 1 db. Vcell fl v 3 b E s (2.7) Mechanical term primary turbulence term elementary subprocesses term The above model brings to the fore the parameters deemed critical in the separation of minerals. The rate of mineral recovery is determined by the mechanical properties of the cell (gas flow rate, bubble size and cell volume), the cell turbulence (expression in brackets in the equation), and the collection efficiency. The number and size of bubbles and particles, the pulp viscosity, and the power input all play an important role in the rate of flotation. In plant operations, when mechanical flotation cell properties and the turbulence are maintained constant, the flotation rate depends on the collection efficiency. The three sub-processes constituting the collection efficiency are further explored below Particle Collection Efficiency An understanding of the underlying sub-processes governing particle collection in the pulp and froth phases is important in process optimisation. Thus, an understanding of bubbleparticle interactions is a prerequisite to interpreting the flotation response of minerals. The overall process of particle collection by bubbles, the collection efficiency, is the product of the collision, attachment and stability efficiencies when the froth recovery factor is assumed to be unity for shallow froth depths in small scale cells (Najafi et al., 2008). In the sections that follow, each of these sub-processes is described briefly to emphasize its role in the particle collection process. 12

35 Collision Efficiency, E c The collision efficiency (E c ) is mainly controlled by the bulk hydrodynamics in the flotation cell and is independent of surface chemistry. The Generalised Sutherland Equation (or GSE) may be used to determine the efficiency of bubble-particle collision under conditions of potential flow and takes account both interception and inertial forces (Dai et al., 1999; Ralston and Dukhin, 1999; Ralston et al., 1999a; Ralston et al., 1999b). A hypothetical schematic representation of a particle of diameter d p approaching and sliding around a bubble of diameter d b is shown in Figure 2.1. Figure 2.1 Schematic representation of a particle of diameter d p sliding around a bubble of diameter d b from a collision or adhesion angle of θ a, to a maximum sliding angle π/2. θ t is the maximum possible collision angle or angle of tangency (Dai et al., 1999). An expression for the collision efficiency, E c, is given by (Dai et al., 1998) E C E C SU sin 2 t e 3 2 cos t cos 3 t 3K3 ln cos E 4 C _ SU sin t t (2.8) where the Sutherland collision efficiency E c-su is given by d EC SU 3 d p b (2.9) and accounts only for interception, where - (2.10) 13

36 - (2.11) where d p and d b are particle and bubble diameters respectively, θ t is the maximum possible collision angle (Figure 2.1) of the particle on the surface of the bubble beyond which collision is prevented, and is a dimensionless number that is a measure of the relative importance of the interceptional and inertial contributions to the collision process and is given by 4E C _ SU 9K 3 (2.12) where K 3 is defined as (2.13) where is the liquid viscosity. It is evident that the bubble and particle sizes, bubble velocity, fluid viscosity and the density difference between fluid ( ) and particles ( ), fl p are all critical in the collision efficiency. Collision efficiency increases with particle density. In the model by Bloom and Heindel, (2002), the collision frequency increases significantly when the particle diameter approaches the bubble diameter and superficial gas velocity. The most significant parameter affecting collision frequency is the bubble diameter. The collision frequency also increases with increasing particle radius because of the rapid increase in particle momentum resulting in a higher level of bubble-particle collisions. An increase in turbulent energy dissipation results in an increase in collision frequency (Bloom and Heindel, 2002) Attachment Efficiency, E a The attachment efficiency can be calculated from the measured collection efficiency when the stability efficiency is unity and a particular collision model is assumed (Pyke et al., 2003). A modified Dobby-Finch attachment model (Dai et al., 1999) may be used to calculate the attachment efficiency (E a ), defined as the ratio of the projected area corresponding to an angle θ a, to the projected area corresponding to θ t. The basic 14

37 mathematical expressions of this model are shown below, where the angle θ is measured from the upper surface of the rising bubble, and θ a is the collision (adhesion) angle where its sliding time is just equal to the induction time if a particle collides with a bubble at this angle (Pyke et al., 2003). The angle θ t is the maximum possible collision angle as shown in Figure 2.1 E a sin sin 2 2 a t (2.14) The adhesion angle is given by (Dai et al., 1999; Duan et al., 2003; Pyke et al., 2003) (2.15) where t ind is the induction time, and and p b are particle and bubble velocities respectively. An expression for the sliding time under potential flow conditions has a maximum sliding angle θ=π/2, and is given by: - (2.16) The Induction Time and Contact Angle The induction time is the minimum time necessary for the attachment of an air bubble to the mineral surface (Ralston et al., 1999a). The water film between a mineral surface and air bubble must thin, rupture and recede before successful attachment can take place. The sum of thin film drainage time and three-phase contact line spreading time constitute the bulk of induction time because the rupturing process is considered to be very fast (Ralston et al., 1999a). An empirical equation for the induction time is given by (2.17) The values of A and B in the induction time equation were shown to be independent of the size of particles (diameter less than 80 µm) and bubbles (diameter between 0.7 and

38 mm), and solution ionic strength (up to 0.1M) (Dai et al., 1999; Duan et al., 2003; Pyke et al., 2003). The value for parameter B was found to be 0.6 ± 0.1. Parameter A was found to be inversely proportional to the contact angle and attachment was unlikely to occur for particle contact angles less than 30 o (Dai et al., 1999). The result suggests that particles across all size fractions float when they have contact angles equal to, or above 30. There is, however, no evidence to suggest that the observation holds true for a variety of minerals with different densities, a subject of this current thesis. The role of the mineral surface hydrophobicity in the attachment efficiency is confined to influencing the induction time. The contact angle is considered only a thermodynamic quantity without a direct bearing on the flotation kinetics. However, a study showed that flotation rate increased when the proportion of hydrophobic xanthate increased on a mineral surface (Goryachev and Nikolaev, 2006). An increase in the hydrophobic species on a mineral surface may correspond to increase on the contact angle of that surface, thus providing an indirect link between the contact angle and flotation kinetics. The induction time measured on a hydrophobic surface was found to be much shorter than that measured on a hydrophilic surface (Najafi et al., 2008). This was thought to be due to the repulsion of water by the hydrophobic surface, making the intervening liquid film between bubble and solid surface unstable, and leading to faster film rupture and shorter induction time; hence the attachment efficiency was higher. A hydrophilic surface was thought to induce a stable water film between the bubble and the solid surface so that additional forces were required to thin this liquid film, leading to longer induction time, and hence poor attachment efficiency (Najafi et al., 2008). A rough, hydrophobic surface shortens the time between collision and attachment of a bubble and particle due to the rapid rupturing of the intervening liquid by the asperities (pillars) on the surface (Krasowska and Małysa, 2005; Krasowska and Malysa, 2007). Further, the rupturing of the wetting film between a bubble and a hydrophobic surface may also be due to nucleation of sub-microscopic gas bubbles present on the surface. The number and size of the bubbles depends on surface roughness (Krasowska et al., 2009). Roughness of hydrophobic mineral surfaces therefore enhances the attachment efficiency by shortening the induction time (Najafi et al., 2008). The increase in the flotation rate with an increase in the area fraction of the hydrophobic component may be attributed to the effect of contact angle on induction time. From these 16

39 findings, it can be concluded that the contact angle is the voltage, as it were, of the flotation rate. It seems reasonable, then, to expect particles of the same size and exhibiting the same contact angle to float at the same rate, assuming the induction time is similar, and the cell hydrodynamics is the same. In this thesis, the dependence of the flotation rate constant on the particle contact angle is explored in detail, under constant hydrodynamic conditions (i.e. constant impeller speed, constant gas flow, and constant bubble size through flotation). Establishing the influence of the particle contact angle on flotation recovery and rate constant may lead to new insights that may revolutionize flotation modelling Stability Efficiency, E s Both the hydrodynamics and the attachment force between the bubble and particle control the stability efficiency (E s ). The probability of bubble-particle aggregate stability depends on the attachment force between the bubble and the particle in relation to the external stress forces in the environment (Pyke et al., 2003). The average mean energy dissipation is used in determining aggregate stability (Newell and Grano, 2007). Mean energy dissipation is an indicator of the degree of turbulence within the cell as discussed below. The force balance for a spherical particle can be given as (Pyke et al., 2003) Fca Fhyd Fb Fg Fd F 0 (2.18) where the forces acting upon the particle are the capillary, hydrostatic, buoyancy, gravitational, machine acceleration and capillary pressure terms, respectively, and are equal to zero at equilibrium. The ratio between attachment and detachment forces determines aggregate stability and is given by a dimensionless parameter analogous to the Bond number (Pyke et al., 2003) B * 0 F F det att F g F F b ca F F d hyd F (2.19) The equation for the Bond number may be derived as (2.20) 17

40 where is the surface tension of liquid, θ is the contact angle, 180 (2.21) 2 and g is the gravitational constant. The stability efficiency is related to the attachment and detachment forces by (2.22) - - (2.23) E S 1 e 1 1 * B O (2.24) Attachment is enhanced by the capillary and hydrostatic forces while gravitation, buoyancy, machine acceleration and capillary pressure in gas bubble contribute to the detachment forces. The above expression may be represented as: - - (2.25) where the centrifugal acceleration, a c, depends on the level of turbulence in the cell. When the particle size is smaller than the bubble size, the acceleration is approximated by 2 3 ac 1.9 1/ 3 d b (2.26) for large turbulent eddies. When eddies and bubble-particle aggregates have similar sizes, the acceleration is approximated by (2.27) The stability efficiency is basically a force balance between the attachment and detachment forces in the flotation cell. When the hydrodynamics is maintained constant, any changes to the stability efficiency can be attributed to the attachment forces, which are influenced by the contact angle. When the fraction of the hydrophobic component on a mineral 18

41 surface increased, the detachment force for an air bubble attached to the surface increased (Goryachev and Nikolaev, 2006). The maximum detachment force was realised with a chemically homogeneous surface, as opposed to one with chemical heterogeneity. For a chemically heterogeneous surface, the flotation rate depends on the effective portion of the hydrophobic component (Goryachev and Nikolaev, 2006). Given that mineral particles are heterogeneous both chemically and physically (Piantadosi et al., 2000; Piantadosi, 2001; Piantadosi and Smart, 2002; Priest et al., 2008), the probability of attachment and detachment to and from an air bubble, respectively, is dependent on the contact angle of the surface that presents itself to the bubble. The balance between the attachment and detachment forces acting on a bubble-particle aggregate may determine the critical contact angle for stable bubble-particle attachment. The critical contact angle will be particle size dependent (Crawford and Ralston, 1988). A single hydrophobic mineral particle of defined size may float with the possibility of a distribution of rate constants depending on the contact angle of the surface and the resulting attachment and stability efficiencies. It is not possible to determine the flotation rate constant of one mineral particle in a conventional flotation cell. A single bubble recovers a number of particles at the same time so that the rate constant for those particles reflects the total water repellency of the bubble-particle aggregate Flotation Kinetics and Floatability Components The flotation process is represented by a chemical kinetic analogy where particles and bubbles react to form the bubble-particle aggregate. The general rate equation for flotation is given by (Polat and Chander, 2000) - (2.28) where C p (t) and C b (t) are the concentrations of the particles and bubbles at time t, respectively. The exponents, m and n, are the respective orders for particles and bubbles and k(t) is a rate constant that depends on various parameters governing the flotation process and may vary with time (Polat and Chander, 2000). Researchers generally agree that the recovery of particles is first order with respect to the particles, provided the bubble concentration remains constant (Sutherland, 1948; Imaizumi and Inoue, 1963; Harris and Chakravarti, 1970; Harris and Cuadros-Paz, 1978). The equation is simplified to 19

42 - (2.29) For a batch flotation test, the recovery of one floatable species is given by a solution of the rate equation R 100(1 e kt ) (2.30) where t is the cumulative flotation time (Ralston, 1992; Rahal et al., 2000). The recovery at time t is related to the maximum recovery at infinite time by (2.31) where R max is the maximum recovery at infinite time or maximum recovery possible, and k is the flotation rate constant of the floatable component. A modified flotation rate constant which takes into account both maximum recovery and rate constant has been proposed (Agar et al., 1986; Sripriya et al., 2003). In this thesis, the modified flotation rate constant is called the undistributed rate constant, k*, and is calculated using: krmax k* 100 (2.32) As has been noted before, the floatability of an ore arises from a range of particle properties such as size, density, shape, mineralogical composition, degree of surface oxidation, type and extent of reagent coverage, and the degree of aggregation of the particles in the system (Runge et al., 2003b). With such an array of parameters affecting floatability, the number of floatability components can be large and difficult to determine as they require a large amount of data. Thus, multi-component floatability behaviour is observed for particles within the same size fraction due to different levels of liberation (Sutherland, 1989), hydrophobicity and size (Imaizumi and Inoue, 1963). The rate constant (k) is related to the floatability (P) by (Gorain et al., 1996; Gorain et al., 1998a; Gorain et al., 1998b) (2.33) where S b is the bubble surface area flux (s -1 ), and R f is the froth recovery factor. The floatability, P, is variable and difficult to control due to the large number of parameters affecting it. The total flotation recovery is often used to estimate the rate constant distributions without giving regard to factors such as particle size and contact 20

43 angle (Polat and Chander, 2000). Various distribution functions to account for the variability of k have been proposed by different investigators. The statistical distributions include triangular, gamma, rectangular, sinusoidal, and normal (Polat and Chander, 2000). None of the distributions adequately fits rate constant data from a wide range of flotation conditions. The use of a preselected function to describe the rate constant distribution is not satisfactory in most cases, most likely because the floatability classes chosen do not reflect the actual properties of the particles (size, and most critically, contact angle). One approach to the determination of the distribution of k is to assume that the feed can be broken into a number of components following first order rate constants, e.g., a two component model incorporating a non-floating component (Jowett, 1974). Another approach to estimate the floatability distribution is to assume that the particle floatabilities are proportional to size distribution (Huber-Panu et al., 1976), or to size distribution and liberation class (King, 1976). Flotation results may be presented as plots of the negative logarithm of the fraction remaining versus time. A straight line on such a plot is an indication of first order flotation behaviour in which the rate constant may be obtained from the slope of the line (Sutherland, 1989). If the flotation results produce non-linear plots, it may be evidence of multi-component flotation arising from differences in floatability for each component. Sutherland (1989) found that particles within a size fraction and in all liberation classes displayed two-component flotation behaviour so that it was not possible to link one component to a certain degree of liberation or particle composition. These results suggest heterogeneity of properties within the same size fraction and within the same liberation class. In this thesis, it is assumed that multi-component flotation behaviour within a size fraction and within any liberation class can be explained in terms of the contact angle. Sutherland assumed that chalcopyrite particles of different sizes had the same degree of hydrophobicity (contact angle). It is possible that particle contact angle varies with size, and even within the same size fraction and within a liberation class. Runge and co-workers used a property based floatability component model to study the mineralogical classes in different streams of an industrial circuit (Runge et al., 2003a). The flotation rate was found to be a strong function of size and the mineral composition of the particle surface, although the two parameters did not adequately explain all the variations observed, possibly due to the use of a cyclosizer which cleans the surface, possibly altering 21

44 the surface composition. The reference to surface species implies the contact angle of the particles, which, together with the size, influence the flotation rate and its distribution. The work of Vianna (2004) further characterised the dependence of the floatability parameter, P, on particle size, collector coverage, and liberation. It was found that particle size, collector coverage, and liberation are important in determining how particles float, and may provide a basis for the development of flotation models. In the work of Sutherland (1989), multi-component flotation behaviour was obtained for each liberation class and each size fraction, further complicating the accurate determination of floatability. It may be that multiple floatability behaviour is determined principally by the contact angle, which is the surface area weighted average of the hydrophilic and hydrophobic domains on a mineral surface. Thus, the degree of liberation within a size fraction may not fully explain multi-component flotation behaviour because particles within the same liberation class may have different contact angles and their flotation response is different. The reference to collector coverage indirectly refers to the contact angle of particles. It has been observed that the separation of particles by flotation conserves the floatability components originally in the feed. Runge and co-workers, (2006) developed the concept of nodal analysis, similar to material balances, in which the floatability components in the feed where found to be equal to the sum of the components in the various process streams of the plant. This concept demonstrates that particle separation behaviour can be predicted based on conservation of floatability (Runge et al., 2006). In this thesis, the conservation of floatability components is demonstrated by direct measurement of contact angles on the feed and products of the flotation process. With regard to characterising flotation response on the basis of particle size and contact angle, Blake and Ralston (1985), Crawford and Ralston (1988), and Gontijo and coworkers (2007) showed that there exists a critical surface coverage, or critical contact angle, for particles of a given size fraction to float. On the other hand, liberation does not have a critical liberation reported in the literature for stable bubble-particle attachment: flotation recovery was significant even with a degree of liberation less than 10% in the work reported by Sutherland (1989), showing that particles float, not necessarily based on their degree of liberation, but on the contact angle. This may be the reason why fast and slow-floating components were obtained even with low liberation class particles. 22

45 The approach that is proposed in this work is to use well-defined size fractions of a single mineral, of known contact angle, in the feed, so that the effect of particle size and hydrophobicity on floatability can be characterised. This approach may be a better predictor of flotation response because it is based on quantifiable parameters of particle size and contact angle. An underlying assumption of the classical first-order rate model is that particles have a single-valued rate constant if the analysis is carried out on fractions that are sufficiently narrow with respect to size and specific gravity (Polat and Chander, 2000). The use of a single mineral ensures that the mineral specific gravity is constant across all size fractions and the flotation rate constant is single-valued. The approach therefore eliminates the need to fit a predetermined distribution on the total recovery without regard to other factors. Under the specific conditions used by Imaizumi and Inoue (1963), the flotation recovery pattern with an initial pulp density of up to 69 g/l of hydrophobic quartz in the size range microns followed first order behaviour. The flotation rate constant was considered to be the best descriptor of floatability (Imaizumi and Inoue, 1963) Wettability and Contact Angle The wetting behaviour of mineral particles is described by the contact angle. A gas bubble will minimise its free energy when it comes into contact with a solid surface or liquid surface. The surface free energy is calculated by the product of the interfacial tension and the interfacial area of contact (Rao, 2004). Energy minimisation involves a decrease in interfacial area,, given by A A SG A SL A LG (2.34) where SG, SL, LG represent solid/gas, solid/liquid and liquid/gas interfaces respectively. The free energy change is given by G SG A SG SL A SL LG A LG (2.35) where, are the surface tensions for the solid/gas, solid/liquid and liquid/gas SG SL, LG interfaces. The free energy change of a unit area of the solid/liquid interface is given by the Dupre equation: G SG ( LG) SL 23 (2.36)

46 The thermodynamic equilibrium is expressed by the Young equation as follows (Fuerstenau and Raghavan, 1976; Fuerstenau and Chander, 1986): SG SL LG cos (2.37) Combination of the Young and Dupre equations gives: G LG(cos 1) (2.38) A negative G means there is spontaneous attachment between the air bubble and the mineral surface. Equation 2.38 is the thermodynamic criterion of flotation and indicates that the free energy of the solid/liquid interface decreases with an increase in the contact angle, θ. For a finite value of the contact angle, there is a free energy decrease upon attachment of a mineral particle to an air bubble (Fuerstenau and Raghavan, 1976). The more negative the free energy change, the greater is the probability of flotation. It is now generally recognized that when other parameters are fixed, the flotation response of minerals is dependent on the particle size and the degree of hydrophobicity (Trahar, 1981a). The advancing contact angle is the contact angle when the three phase contact (TPC) line moves over a surface, replacing the gas phase, while the receding contact angle is when the TPC line moves over a pre-wetted surface, thus pushing away the liquid phase. The movement of the TPC line is often impeded by barriers such as surface roughness or chemical heterogeneity (Nadkarni and Garoff, 1992). In the case of water, hydrophobic domains on the surface will pin the motion of the TPC line, resulting in an increase in the contact angle. When the water recedes, hydrophilic domains hold back the draining motion of the TPC line, thus decreasing the contact angle. From this analysis, advancing contact angles characterize the degree of hydrophobicity while the receding contact angles are an indication of the degree of hydrophilicity. With this in mind, it would seem to make sense to relate flotation response to a measure of surface hydrophobicity (advancing contact angle). The basis for mineral separation is the response of mineral surfaces to wetting liquids, normally water. The use of the advancing contact angle of particles may give an appropriate measure of particle hydrophobicity which may then be correlated to flotation response. The advancing contact angle measured on particles compared more favourably with flotation recovery than that measured on polished surfaces (Subrahmanyam et al., 1999). Some researchers (Woods et al., 1993; Kelebek and Yoruk, 2002) have found poor 24

47 or no correlation between the contact angle and flotation recovery, probably because of the differences in techniques used in measuring the contact angle and the material surface under consideration. This is discussed further below. With regard to bubble-particle attachment, the receding contact angle, θ R, is important while the advancing contact angle, θ A, is relevant to bubble-particle aggregate stability (Hornsby and Leja, 1983). From this analysis, a low receding contact angle shows resistance to de-wetting, and is not good for the attachment process. The water advancing contact angle, θ A, is always larger or equal to the receding contact angle, θ R, and the difference between the two is called contact angle hysteresis, which is caused by chemical and/or physical heterogeneity of the surface (Adam, 1963; Gaudin et al., 1963; Tomlinson and Fleming, 1963a; Nadkarni and Garoff, 1992; Chibowski et al., 2002). Despite the existence of heterogeneities in particle shape and roughness, particles of the same mineral and of a particular size fraction may give the same flotation response (maximum recovery and rate constant) over a certain narrow range of contact angles. If this were the case, the flotation response can be characterized in terms of particle size and contact angle, over a limited range of each parameter Hydrophobicity and Sulphide Mineral Surface Characteristics The efficient separation of minerals by flotation takes advantage of the different surface species existing in the pulp between the value mineral and the gangue. The degree of hydrophobicity depends on the amount and type of species on a mineral surface (Piantadosi and Smart, 2002). Work by Piantadosi and Smart (2001) demonstrated that mapping of hydrophobic and hydrophilic species on mineral surfaces was possible using ToF-SIMS and the results could be correlated to surface hydrophobicity. Surface chemistry is the principal determinant of the contact angle (Urquhart et al., 2008), which in turn affects the bubble-particle attachment and collection efficiencies consecutively, and subsequently, the flotation rate constant (Hart et al., 2006). Research into mineral surface reactions and the mechanisms involved has been enhanced by the development of surface analytical techniques such as X-ray photoelectron spectroscopy (XPS) and time of flight secondary ion mass spectroscopy (ToF-SIMS), (Mielczarski et al., 1996a; Laajalehto et al., 1997; Khmeleva et al., 2005; Hart et al., 2006). In the case of chalcopyrite, initial reactions on the mineral surface in an alkaline solution are thought to involve the oxidation of the iron to oxides and hydroxides (Fe 2 O 3 /Fe(OH) 3 ) 25

48 forming a monolayer (Vaughan et al., 1997). Sulphur enrichment on the surface takes place when chalcopyrite is held in aqueous solution without a collector (Mielczarski et al., 1996a). A metastable phase of CuS 2 is formed in the process, with the possibility of further decomposition by subsequent reactions producing copper oxides and hydroxides on the surface. When chalcopyrite is conditioned with xanthate collector, common species observed are adsorbed copper xanthate, dixanthogen and oxidation products, with the likelihood of iron xanthate (Mielczarski et al., 1996a). The amount of each species is dependent on the ph and the concentration of the xanthate solution (Mielczarski et al., 1996b). The hydrophobic copper xanthate is believed to form simultaneously with iron hydroxide species, forming an electrochemical redox couple with the reduction of oxygen as the cathodic reaction. The formation of dixanthogen, which is the oxidation (dimerization) of the collector ions (X - ) may also take place in the pulp according to the reaction (Smart et al., 1998) 2X X 2 2e (2.39) further adding to the hydrophobic species available. Dixanthogen is very unstable and decomposes when the mineral surface is dried, so that no dixanthogen is observed on surfaces of dry samples. There is still some uncertainty as to the exact manner in which collector molecules and mineral surfaces interact. There are two mechanisms put forward in the formation of the copper xanthate complex. The first one is the chemisorption of xanthate (MX 2 ) ion to copper that is in the mineral surface producing a monolayer, and this is thought to be responsible for floatability (Goh et al., 2008). The second is the precipitation of the metal salt, copper xanthate (CuX ads ), onto the mineral surface thereby producing a hydrophobic multilayer (Prestidge and Ralston, 1996a). However, Goh et al. (2008) expressed caution about the second mechanism, arguing that most of the floatability is due mainly to xanthate chemisorbed to copper atoms in the surface. In the pulp, any of the five species, i.e., X ads, MX 2, S, X 2 and metal deficient sulphide, may contribute to the hydrophobicity of the mineral surface (Smart et al., 1998). The reactions considered for chalcopyrite are representative of the mechanisms followed by iron containing sulphide minerals such as pyrite, pyrrhotite, chalcopyrite and pentlandite. Surface products and mechanisms may include any of the following (Smart et al., 1998): metal deficient (sulphur-rich) oxide surfaces, polysulphides and elemental sulphur; 26

49 oxidized fine particles attached to larger sulphide particle surfaces; colloidal metal hydroxide particles and flocs; continuous surface layers (e.g., oxide/hydroxide) of varying depth; formation of sulphate and carbonate species; non-uniform spatial distribution with different oxidation rates, e.g., isolated, patch wise oxidation sites that are face specific. Mineral surfaces are usually heterogeneous so that the overall hydrophobicity and flotation response will be controlled by the arrangement of hydrophilic and hydrophobic patches (Fairthorne et al., 1997; Smart et al., 1998; Hart et al., 2006). Mineral particles have different adsorption densities of collector, leading to variation in surface hydrophobicity (Song et al., 2001). In flotation, this leads to different flotation rates and floatabilities even for particles with the same size and density. Piantadosi (2001) also found that a large variation in collector concentration on particles of the same size existed, which confirms observations by Song et al. (2001). The possible reasons given were that this is a result of non-uniform distribution of active sites and their degree of activity on mineral particles. The variation of collector concentration from particle to particle may lead to different contact angles on the surfaces of individual particles and a distribution of contact angles in a sample. It is doubtful that all sulphide mineral particles have the same contact angle in a powder although subjected to the same chemical treatment and preparation, as suggested in tests on methylated quartz (Blake and Ralston, 1985; Crawford and Ralston, 1988; Gontijo et al., 2007). Evidence of apparent contact angle heterogeneity within a size fraction is suggested by the presence of a non-floating fraction, which reports to the tailing, while some particles, even of the same size fraction, report to the concentrate by genuine bubbleparticle attachment. Direct evidence for contact angle heterogeneity on single size fractions of a chalcopyrite sample has recently been provided by ToF-SIMS examination on individual particles (Brito E Abreu, 2010). In the latter study it was found the contact angle on single mineral chalcopyrite within a narrow size fraction and similarly prepared, varied from particle to particle, approximating normal or gamma distribution functions (Figure 2.2) depending on the mean contact angle value. 27

50 Figure 2.2 Histogram of feed contact angle distribution from ToF-SIMS (a) µm Chalcopyrite sample conditioned with 2g/t DTP, CA=55, (b) µm Chalcopyrite sample conditioned with 2g/t DTP, CA=60. (Brito E Abreu, 2010) The average contact angle of a chemically heterogeneous but smooth surface may be approximated by the Cassie equation as discussed below Cassie Equation The equilibrium contact angle measured on a chemically heterogeneous but smooth surface is described as (Cassie and Baxter, 1944) (2.40) 28

51 where θ c is the measured contact angle, f i is the surface area fraction of a chemical domain with contact angle θ i, and f i = 1. The equation is based on the assumption that the droplet is large compared to the surface area fractions, which are larger than the molecular scale. The Cassie equation has been used to calculate the contact angle of heterogeneous mixtures of hydrophilic and hydrophobic particles (Stevens, 2005; Priest et al., 2008) Influence of Collector Chain Length on Contact Angle The role of a collector is to impart hydrophobicity to the surface of a mineral. The contact angle of the surface reaches a maximum as the collector concentration is increased. This maximum contact angle is characteristic of the collector used, and not the mineral substrate. The characteristic contact angle increases with the alkyl chain length (Gupta and Yan, 2006; Somasundaran and Wang, 2006). Contact angle values determined for various collector chain lengths (Table 2.1) are in good agreement with those measured on polished mineral surfaces, regardless of the type of mineral substrate (Sutherland and Wark, 1955). These values may differ from those measured on powdered particles due to particle geometry and surface roughness effects. Table 2.1 and Wang, 2006) Calculated values of contact angles (θ) of n-alkyl groups (Somasundaran Contact Angle and Mineral Surface Roughness Mineral surface roughness is one of the causes of contact angle hysteresis. Large contact angle hysteresis can be of advantage to the recovery of hydrophobic minerals. The detachment force for a bubble-particle aggregate increases with contact angle hysteresis 29

52 (Schimann, 2004). A rough surface may therefore have a positive effect on the stability efficiency (Sutherland and Wark, 1955) Wenzel Equation The effect of surface roughness on the contact angle of a surface can be described by the Wenzel equation, (Wenzel, 1936) (2.41) where θ w is the contact angle on the rough surface, r w is a roughness ratio of the actual surface area to the apparent surface area, and θ is the contact angle on a chemically identical smooth surface. Roughness increases the contact angles of hydrophobic surfaces (θ>90 ), while it decreases the contact angle of hydrophilic surfaces (Miller et al., 1996; Extrand, 2004; Birdi, 2009). If the asperities are sharp or needle-like, they significantly increase the contact angle (Nakajima et al., 2001). Mineral particles are often rough with sharp edges that may increase or decrease the contact angle of hydrophobic or hydrophilic particles respectively, above or below that measured on planar surfaces. The advancing contact angles measured by the Washburn or Equilibrium Capillary Pressure techniques take into account particle shape and roughness by virtue of the measurement method in which testing liquids wet whole multiple particle surfaces as discussed further below Critical Contact Angle and Recovery The recovery of minerals increases with an increase in contact angle (Blake and Ralston, 1985; Prestidge and Ralston, 1996b). As noted before, a critical contact angle for floatability of mineral particles of different sizes exists (Blake and Ralston, 1985; Crawford, 1986; Gontijo et al., 2007). Below this critical angle, mineral particles do not float under any condition, and report to the tails. For particle sizes above 20 microns, the critical contact angle (Figure 2.3) apparently increases with particle size (Crawford and Ralston, 1988), with fine particles (5-14 µm) requiring a critical contact angle of about 60 (Crawford and Ralston, 1988; Miettinen, 2006). 30

53 Figure 2.3 Particle size as a function of critical surface coverage and contact angle. Methylated quartz particles. (Crawford, 1986) Although the results reported recently by Gontijo et al. (2007) cover a much larger size range, with a maximum of 1000 µm, the critical contact angles reported for hydrophobised quartz do not show a clear trend with size (Figure 2.4). This might be due to the method used in the determination of the critical contact angle for each particle size. The point of intersection between the contact angle axis and a line drawn through the point of inflection of a recovery versus contact angle curve was considered the critical contact angle. The authors do not explain the physical basis of this approach, which has the disadvantage of underestimating or overestimating the critical contact angle of a particular size fraction due to possible heterogeneity of contact angles within a sample of particles of the same size. 31

54 Figure 2.4 Critical contact angle determination. Quartz recovery after 8 minutes of flotation, obtained in a flotation column. (Gontijo et al., 2007) Critical contact angles of 1 o for small size particles of low density have been reported, and some large particles require up to 90 o contact angle before they can float (Hornsby and Leja, 1983). Theoretical analyses of maximum floatable particle size indicate a minimum contact angle of about 10 o to 20 o for particles above 50 µm in size, for dynamic conditions in a flotation cell (Ralston and Newcombe, 1992). These results show that the critical contact angle is important in determining floatability, but the method for its determination has not been developed to a stage where there is a standard procedure. It is expected that factors such as mineral specific gravity, particle size and hydrodynamic conditions play an important role in determining the critical contact angle for stable bubble particle attachment. It would appear that the non-floating component in each size fraction has a contact angle at or less than the critical required for stable bubble-particle attachment (Crawford et al., 1987; Crawford and Ralston, 1988; Gontijo et al., 2007). A critical contact angle appears 32

55 for all size fractions, with the fine and coarse size fractions having apparently greater values than intermediate size fractions (Crawford, 1986; Miettinen et al., 2000; Gontijo et al., 2007). If the distribution of contact angles within a sample of a given size fraction can be approximated with a statistical function, it may be possible to estimate the number of particles with contact angle less than the critical value. When the critical contact angle of particles is known, the maximum recovery can be predicted from knowledge of the feed contact angle distribution. This approach to flotation analysis is explored in this thesis for the first time and is expanded upon on other mineral-collector systems and a natural ore Contact Angle Measurement Methods The contact angle can be measured on flat/polished mineral surfaces or on packed mineral powder beds (Dunstan and White, 1986; Ralston and Newcombe, 1992). The former technique gives specific surface information while the latter gives the average particle wettability for a particular powder bed (Stevens, 2005). The quantitative measure of wettability, the contact angle, of finely divided powder material is of interest to the flotation process. Measurement of contact angles on flat surfaces (sessile drop technique) makes use of digital imaging techniques to capture the bubble or droplet image on a solid surface and the image is processed and used to determine the contact angle. Mettinen (2006) used the film flotation technique to measure the contact angle of fine particles. The method is based on the critical surface tensions determined by use of methanol solutions. The advancing water contact angles of these particles were estimated from a calibration curve of advancing water contact angle and critical surface tension. Results obtained using this technique were in agreement with those reported for identically treated surfaces (Marmur et al., 1986). Another approach to measure contact angles for particles is to attach the particles to a flat plate and measure the contact angle of the plate using the Wilhelmy balance. Bröckel and Löffler (1991) measured the contact angle of glass particles from 40 μm to 1 mm using this technique. An inert adhesive of known contact angle was used to adhere the particles. This method is sensitive to the porosity of the samples (Brockel and Loffer, 1991), but reproducibility is a major challenge with this technique because of variable surface loadings of particles on the plate surface. 33

56 The Washburn and Equilibrium Capillary Pressure (ECP) techniques are used to measure contact angles on powder particles (Stevens, 2005). Both methods make use of capillary pressure to drive a liquid at an observable rate through a packed bed of particles in a capillary tube. The Washburn method measures the advancing contact angle only while the ECP technique measures both the advancing and receding contact angles. The methods may have problems in systems where surface-active agents are present. If there is dissolution of surfactant this will alter the initial contact angle, or if the liquid used has some dissolved surfactants, these will adsorb on particle surfaces on the wetting front, thus depleting surfactant molecules (Ralston and Newcombe, 1992). Other issues with the Washburn technique include reproducibility of the packing of particles in a capillary tube (Chau, 2009b), an aspect that will be tested in this thesis. The Washburn and ECP techniques measure the average contact angle of a bed of particles, and not the individual particle hydrophobicity. Other limitations also relate to the maximum particle size for accurate and reliable measurements, which is not clearly defined in the literature. The Washburn technique is limited to a maximum contact angle of 90. Despite these limitations, the Washburn technique is cheap and easy to use. Measurements can be carried out in a short time, and the contact angles thus measured take into account the particle shape and roughness. The Washburn and ECP techniques are the best available methods to measure the contact angle of particles at the moment, and were used in this study. More details are given in Chapter 3 (Section 3.2.2) as the theory and method descriptions are more relevant there. Attempts are under way to develop a method of inferring surface wettability by use of surface analysis techniques such as the Time of Flight Secondary Ion Mass Spectroscopy (ToF-SIMS). The approach correlates the advancing contact angle measured using the Washburn technique to the surface species intensities. The method involves the simultaneous measurement of the Washburn advancing contact angle and carrying out surface analysis using ToF-SIMS on particles of the same size fraction and subjected to the same treatment. Relevant statistical methods are used to determine the principal species responsible for particle hydrophobicity, resulting in the generation of an empirical relationship between the advancing contact angle and surface species intensities. This was trialled in the work of Stevens (2005) and more recently by Brito E Abreu et al. (2010). In the latter work, the empirical relationship between the contact angle and surface species was developed using single mineral chalcopyrite conditioned with sodium dicresyl 34

57 dithiophosphate (DTP) (Section ). The ToF-SIMS approach is useful, but has the disadvantage that the method requires a time consuming and expensive ex-situ process. Rather, the approach discussed in this thesis is based on the flotation method itself, and eliminates the need for drying of particles once it is fully developed Hydrodynamics The parameters that determine flotation cell hydrodynamics have been largely kept constant through the flotation process and in different tests in this current work by fixing the agitation speed, gas flow rate, initial pulp density, volume, and frother concentration for each set of tests. However, it is important to have an understanding of the impact of each of these parameters on the flotation process itself. What follows is a brief review of the important hydrodynamic aspects of flotation Power Input and Energy Dissipation For the efficient operation of a flotation cell, the pulp needs to be agitated or energy input is required to increase the momentum of particles. Flotation cell hydrodynamics (macroscopic flow of fluid in a cell) is largely dependent on the impeller type and speed and on the cell characteristics such as size and shape (Deglon, 2005). The hydrodynamics within a cell have been described in terms of parameters such as power intensity, impeller tip speed, Power number, Froude number, tank turn-over time, air flow number, etc (Deglon, 2005; Grano, 2006). The fluid that leaves the impeller zone carries kinetic energy that eventually decays into turbulence (turbulence energy dissipation). Turbulence is composed of physical vortices or eddies of various sizes and time scales, and is often presented as root mean square (RMS) turbulent velocity, turbulent kinetic energy or turbulent energy dissipation (Deglon, 2005). Typical impeller tip speeds range between 4.9 and 8.8 m/s but the majority of plant flotation cells operate in the range of 5.0 to 7.0 m/s. The impeller tip speed affects the velocity at which the pulp leaves the impeller, and influences both gas dispersion and pulp circulation (Deglon et al., 2000). Agitation of the flotation pulp is required so that particles are suspended and this increases the probability of collision between particles and bubbles. As the agitation speed is increased, the flotation rate also increases to a maximum before it decreases due to excessive turbulence leading to detachment of particles from bubbles (Figure 2.5). The 35

58 point at which this happens is a function of particle size, hydrophobicity, density, and possibly the flotation cell impeller design. Figure 2.5 Relationship between flotation rate constant k and impeller speed. (Gupta and Yan, 2006). It is still unclear whether an increase in energy dissipation has any effect on bubble-particle attachment, though it is known that it reduces bubble velocity to a critical value (Newell, 2006b). Increased energy dissipation may affect bubble-particle contact time, and hence the attachment efficiency. Depending on the physical conditions for a particular set up, there is an optimum power input that maximises recoveries, beyond which there will be detachment of particles from bubbles, especially the coarse size fraction. Deglon and O Connor (1997) studied the hydrodynamics of fine particle flotation. They found that the rate of energy dissipation in the impeller zone (60-70% of total input) was orders of magnitude higher than in the bulk of the cell. The impeller zone controls the process of bubble breakup in the cell due to the high energy turbulent eddies. It was also observed that the flotation rate constant increased with increasing power input for both particle sizes used (3 and 24 micron particles) (Deglon and O ' connor, 1997). When frits that produced large bubbles were used, the bubble size decreased with increasing power input, but, when small bubbles were used, there was no observable change in bubble diameter with power input. The frother concentration used was 100 ppm methyl isobutyl carbinol (MIBC), a concentration higher than that reported by other researchers for the 36

59 critical concentration to prevent bubble coalescence (Cho and Laskowski, 2002; Newell, 2006b). An increase in agitation rates increased fine particle recovery but excessive rates resulted in a sharp decrease in grades due to increase in unselective recovery of the -25 micron size particles (Harris et al., 1997). Specific power input decreased drastically as the gas velocity in the cell increased (Grau, 2006a). This could be a result of the higher gas hold-up and reduced pulp density. Since the power input increases with agitation speed the reduced specific power input with an increase in gas velocity suggests that optimum conditions for the two parameters are inversely related. A high agitation speed may be used with a low gas velocity or vice versa. The mean energy dissipation increases when air is sparged into the cell and increases with an increase in superficial gas velocity (Newell, 2006b) Air Flow Rate The amount of air supplied to a flotation cell determines the volume of air and the number of bubbles passing through the pulp. The gas hold-up is the ratio of the volume occupied by air to total volume of the pulp (Newell and Grano, 2007). An increase of the amount of air in the pulp decreases the net pulp density and may result in the settling of coarse particles and subsequent decrease in their recovery. However, the rate of recovery of hydrophobic particles increases with an increase in the number of suitably sized air bubbles in the pulp. Not only is the number of bubbles important, but their size (and possibly distribution) (Kracht et al., 2005) which determines the rise velocity. The effectiveness of air dispersion depends on cell dimensions and design, frother concentration, gas flow rate and impeller rotational speed Superficial Gas Velocity, Bubble Size, and Bubble Surface Area Flux The superficial gas velocity is the volume of gas passing a unit cross-sectional area of a floatation cell in the pulp per unit time (Vera et al., 1999; Rao, 2004; Kracht et al., 2005) Q J g (2.42) A where J g = superficial gas velocity, Q = volume of air flowing per unit time, and A = crosssectional area of the cell. 37

60 An increase in the superficial gas velocity results in an increase in bubble velocity (Newell, 2006b). The bubble velocity decreases with an increase in mean energy dissipation (increase in impeller rotational speed), while it increases with an increase in cell volume. It has also been observed that an increase in the superficial gas velocity results in the coarsening of the bubble size distribution (Kracht et al., 2005; Grau, 2006a) even in noncoalescing systems. The gas hold-up increases with an increase in the superficial gas velocity, J g, and impeller rotational speed (Newell, 2006b). Flotation recovery increases with small bubbles that carry more mineral particles per unit volume of air. The collision frequency and efficiency both increase with a decrease in bubble size. Large bubbles may be a result of poor shearing by the impeller, a high J g and inadequate frother concentration. The average bubble size can be represented by the Sauter mean diameter (d 32 ) or the arithmetic mean bubble diameter (d mean ) (Barigou and Greaves, 1992) d d 32 mean d d 3 i 2 i n d i (2.43) (2.44) where d i is the individual bubble diameter and n is the total number of bubbles. As the impeller speed is increased, the bubble size distributions become narrower and finer (Kracht et al., 2005; Grau, 2006a). At frother concentrations exceeding the critical coalescence concentration (CCC), the maximum stable bubble diameter increased with increasing chain length of the frother molecule. Grau (2006) also found that in noncoalescing conditions, the bubble size increased with an increase in airflow rate, the trends observed being frother specific. The Sauter mean bubble diameter increased with an increase in the percentage of nonfloatable solids in the pulp (Grau, 2006a). The effect was more pronounced at solids concentrations above 20%. This was thought to be due to the damping effect of solids on the turbulent eddies. The bubble surface area flux is a machine parameter that combines J g and d 32. It is a measure of the bubble surface area rising up through the cell per unit cross sectional area, and is given by: 38

61 S b 6J d g 32 (2.45) where S b = bubble surface area flux, J g = superficial gas velocity, d 32 = Sauter mean bubble diameter. Cell performance may improve with an increase in S b, the bubble surface area flux (Finch and Dobby, 1990). A linear relationship exists between the bubble surface area flux and the flotation rate constant as shown by equation The rate constant, the bubble surface area flux and collection efficiency are related by (Jameson et al., 1997) k 3J 2d g b E coll 1 4 S b E coll (2.46) 2.8. Particle Size A typical recovery profile (Figure 2.6) of fine, intermediate and coarse particles is one where there is a low recovery of fine particles, and the recovery increases with size and reaches a peak at the intermediate size range before it decreases for the coarse particles (Pease et al., 2006). Figure 2.6 Conventional recovery profile with particle size (Pease et al., 2006). Results of batch flotation experiments for different minerals show a linear relationship between the rate constant and particle size (Trahar, 1981a). However, these results did not cover the whole size range normally encountered in flotation. The typical variation of rate 39

62 constant with particle size is one where the rate constant increases with particle size, reaching a maximum at a particular size fraction, dependant probably, on contact angle, particle composition, mineral specific gravity and particle shape (Figure 2.7). Lower flotation rates are reported in the fine and coarse fractions and higher flotation rates in the intermediate size range (Castro and Mayta, 1994; Koh et al., 2009). Bubble-particle collision frequency is lower for fine particles and the stability efficiency is lower for coarse particles, resulting in lower recoveries (Pyke et al., 2003). Apparently, angular (Figure 2.7) particles float faster than spherical particles, assuming they have the same contact angle, a result that might suggest that particle shape may significantly alter flotation response. However, the contact angles of the particles are not indicated and it is not clear whether they calculated the distributed or undistributed rate constants, therefore it is difficult to draw a conclusion from the result explicitly, although asperities on the mineral surface may have a positive effect on the induction time as discussed before (Nakajima et al., 2001). Figure 2.7 Comparison of flotation rate constants for ground and spherical ballotini at various sizes (Koh et al., 2009). The effect of particle size on the collision, attachment and stability efficiencies is shown for chalcopyrite particles assumed to have the same contact angle under constant hydrodynamics (Figure 2.8). The collision efficiency increases with particle size, while the attachment efficiency decreases as the particle size increases. In the example in Figure 2.8, the stability efficiency is nearly constant for particles less than about 60 microns and decreases as the particle size increases above 60 microns (Newell, 2006a). The stability efficiency is also influenced by energy dissipation, with stability of the bubble-particle 40

63 aggregate decreasing with an increase in energy dissipation for coarse particles (Deglon, 2005; Newell, 2006). Figure 2.8 Collision efficiency (E c ), attachment efficiency (E a ) and stability efficiency (E s ) of chalcopyrite particles as a function of particle diameter. Contact angle 71 o, stirring speed 720 rpm, d b = 0.97 mm (Duan et al, 2003) Frothers and the Froth Phase in Flotation Frothers are added to stabilize the froth and enhance generation of fine bubbles (Cho and Laskowski, 2002). The size of bubbles for a specific air sparger depends on frother concentration and the agitation rate. Studies that have been made on frothing action have shown that bubble size and distribution are affected by the processes of bubble break up and coalescence. A critical concentration of frother has been defined as one at which bubble coalescence is effectively prevented and the bubble size remains relatively constant with increasing frother concentration and agitation speed (Cho and Laskowski, 2002). At the critical concentration for coalescence (CCC), the coalescence time might be longer than the contact time between bubbles, thus preventing coalescence. The film between colliding bubbles becomes elastic enough to resist rupture (Grau, 2006a). Gourram-Badri et al (1997) found that a concentration of methyl isobutyl carbinol (MIBC) of 30 ppm reduced bubble coalescence to 0%. However, Cho and Laskowski, (2002) report a critical concentration of MIBC of 80 ppm. The difference could be due to differences in the experimental set-up and bubble generation methods. Newell (2006) found that at a frother (MIBC) concentration greater than 20 ppm, the bubble size remained fairly 41

64 constant. Newell s value is much closer to the one reported by Gourram-Badri et al (1997). It appears from these investigations that agitation plays a significant role in bubble break up and coalescence. To fix the bubble size at a certain airflow rate and agitation speed, the critical coalescence concentration should be determined for the type of frother. The Sauter mean bubble diameter decreases with an increase in the frother concentration, reaching a constant value at the CCC (Figure 2.9). Figure 2.9 Effect of frother concentration on bubble size (schematic) (Cho and Laskowski, 2002). Hadler et al (2005) investigated frother-collector interactions and found that about 20% of frother was lost instantly on mixing and that the amount of frother lost was independent of the concentration of collector (SIBX) in solution. The frother was thought to adsorb on mineral surface together with the collector. Without collector, no loss of frother was observed when mineral particles were added (Hadler et al., 2005). It is also thought that frother type may also affect water recovery in the flotation process (Finch et al., 2006). The froth height was found to be proportional to the concentration of MIBC in the pulp (Gourram-Badri et al., 1997). A zinc concentrate (8 % solids) gave a froth height less than that obtained without particles but the height was more stable than that of a copper concentrate that was more hydrophobic. They inferred that there was preferential adsorption of frother on less hydrophobic surfaces (zinc concentrate) than on the more hydrophobic copper concentrate based on these results. The froth height with the copper concentrate was higher than that obtained without solids in the water. Finer particles gave higher froth heights than coarser particles for both concentrates (Gourram-Badri et al., 42

65 1997). Thus, the degree of hydrophobicity, particle size and concentration of floatable particles in the pulp may affect froth characteristics, necessitating adjustments when these parameters change significantly The Froth Phase A flotation cell operates in two distinct phases or zones (Figure 2.10); namely the pulp or collection zone and the froth phase (Feteris et al., 1987; Vera et al., 1999; Seaman et al., 2004). The pulp zone is where the formation of hydrophobic particle bubble aggregates takes place, while in the froth phase there is separation of these aggregates from surrounding suspended hydrophobic and hydrophilic particles. Thus, the froth phase determines the grade in flotation (Seaman et al., 2004). A measure of froth zone performance used extensively in flotation research is the Froth Recovery, Rf, of a particular system. Froth recovery is defined as the mass of particles collected in the concentrate launder by true flotation as a fraction of the mass of particles entering the froth phase (Vera et al., 1999; Seaman et al., 2004). Figure 2.10 Two-phase model description of flotation after (Feteris et al., 1987). Froth zone recovery can be calculated using the ratio between the collection zone rate constant and the overall rate constant (Vera et al., 1999). The flotation rate constant at zero froth depth is the collection rate of the pulp zone (k c ), which is independent of the froth depth. Thus, at zero froth depth, R f, by definition is equal to 100%. At all froth depths greater than zero, froth zone recovery will be less than or equal to 100% (Vera et al., 1999). Thus, the influence of the froth phase is minimal when the flotation process is carried out at shallow froth depths. For strongly floating minerals like chalcopyrite, the froth zone recovery is taken as unity, but this would not be so for weakly floating minerals like the platinum group minerals. 43

66 Froth depth may also affect the rate of true flotation over a range of operating conditions (Ross, 1991). The change in the proportions of hydrophobic and hydrophilic particles in the pulp results in changes in froth stability and overall water recovery (Yekeler and Sönmez, 1997; Hadler et al., 2005). In this thesis, the froth depth is kept constant, and changes to froth stability due to depletion of hydrophobic particles over time is compensated for by the stage addition of frother Summary The literature presented illustrates the effects on the flotation process of the sub-processes of collision, attachment, and stability. The influence of the induction time on the attachment process and its dependence on the contact angle and surface roughness was reviewed. Particle surface hydrophobicity and wetting characteristics are defined by the contact angle. Methods for direct measurement of contact angle of particles have been reviewed. The methods currently available to measure the contact angles of particles cannot be applied directly in a process stream, hence the need for the mineral flotation response approach to infer an operational contact angle of hydrophobic particles. Flotation cell hydrodynamics plays an important role in the recovery of minerals. A summary is presented below with hypotheses to be tested. Mineral hydrophobicity plays an important role in bubble-particle attachment and stability. The higher the contact angle, the shorter the induction time, and the stronger is the bubbleparticle attachment strength. Mineral surface roughness enhances the attachment and stability of hydrophobic bubble-particle aggregates. Mineral recovery increases with an increase in the degree of hydrophobicity. The recovery is also a function of particle size, so that a certain minimum contact angle is required for a particle of a given size to attach to a bubble and remain stable. The degree of hydrophobicity is determined by the amount and type of species on the surface of the mineral and the surface roughness. The method of measuring the contact angle is important. The Washburn approach takes into account surface heterogeneity and roughness, effectively measuring the wettability of the particle bed. Particle collection from the pulp zone of the flotation cell is made possible by air bubbles that collide with hydrophobic particles and form a bubble-particle aggregate. Hydrophobic particle recovery from the pulp is a first order process with respect to particle concentration in the pulp when the bubble surface is not overloaded and the bubble size and 44

67 concentration remain constant. The rate constant is influenced by the contact angle under constant hydrodynamic conditions. Particles float with a distribution of rate constants due to differences in size, composition, shape, roughness and surface chemistry, etc. Differences in surface chemistry, roughness, and particle size give rise to multiple floatability components for a feed sample of particles similarly treated. However, a narrowly sized sample of similarly treated particles may have a single-valued rate constant, assuming similar wetting behaviour. The influence of the froth phase on the rate of particle recovery is minimal when the flotation process is carried out at shallow froth depths. To develop an ore floatability characterization tool, the flotation conditions need to be clearly defined in terms of the gas dispersion parameters, the power input, reagent regime and solids concentration in the pulp. A number of hypotheses are proposed to investigate flotation response based on particle size and contact angle, with all other parameters well defined and kept constant where possible. The following hypotheses have been tested in this thesis: Particle size fractions attain different contact angles when conditioned with collector in the same pulp. Particle flotation behaviour is related to the contact angle achieved, Hydrophobic mineral size fractions float independently of each other so that the first order kinetic model can be applied in the calculation of the flotation rate constant. This also allows the collation of flotation responses based on particle size fraction and contact angle ranges, which leads to the development of calibration curves that can be used to benchmark flotation response of minerals under the specific physical conditions used in this work, There is conservation of floatability components through flotation as demonstrated by Washburn advancing contact angle measurements on flotation feed and products, The maximum recovery of particles in a given size fraction and contact angle range can be modelled assuming a statistical distribution of individual particle contact angle around the mean. This leads to the determination of the critical contact angle for stable bubble-particle attachment, 45

68 When the conditions for first order kinetics are met with respect to hydrophobic particle concentration in the pulp, the influence of pulp density of a non-interacting gangue mineral on the flotation rate constant is insignificant, so that flotation behaviour at 2% solids is the same as that at 30% solids, the bulk being noninteracting gangue, The flotation response of mineral particles of the same type, and within the same size fraction and contact angle range is the same irrespective of the hydrophobisation route, or type of collector used. It is possible to infer an operational contact angle of hydrophobic mineral particles in an ore by benchmarking against the calibration which can be confirmed by direct and indirect measurement of the contact angle. 46

69 CHAPTER THREE 3. Experimental Procedures A general outline of the experimental approach included material preparation, contact angle manipulation, contact angle measurement, flotation, sizing, and quantification of flotation products (Figure 3.1). Figure 3.1 Experimental outline for single mineral work Materials Sample Preparation (Single minerals) Two sulphide single minerals were used in this work, chalcopyrite and pyrite, which have densities of 4.2 and 5.0 g/cm 3, respectively. Chalcopyrite was supplied by Mannum Minerals, SA, while pyrite was supplied by Geo Discoveries (The Willyama Group, 47

70 NSW). Both minerals were supplied in lump form and were crushed to less than 2.4 mm and separately blended. The chalcopyrite, as supplied, was significantly contaminated with quartz (8.5% SiO 2 un-sized assay), while the pyrite was relatively pure, with less than 1% SiO 2 contamination (Table 3.1). A 2 kg single mineral sample (maximum size 2.4 mm) was ground for 15 minutes in a laboratory stainless steel rod mill (10 stainless steel rods with a total weight of 9.3 kg), and 1 dm 3 water at natural ph. The ground product was screened with a 420 µm sieve to remove oversize material. The chalcopyrite was immediately floated without collector in a cleaning stage to reduce the amount of impurities, notably silica, to low levels (Table 3.2). Polypropylene glycol (PPG425) was used as frother. A 5% aqueous solution of the PPG425 was prepared. One molar analytical grade sodium hydroxide was used to adjust the ph to 10. The rejection of silica at this stage relied on the natural hydrophobicity of the freshly prepared chalcopyrite particles. This was important to ensure that multi-component flotation behaviour was not due to hydrophilic impurities such as quartz. Concentrates from this stage were sieved, filtered, dried, stored in sealed bottles, and used as feed to subsequent experiments. Pyrite did not require a cleaning stage and was separated into size fractions after grinding. In order to study the flotation response of the model particles based on size and contact angle, the feed material needed to be of high purity, and separated into well defined metallurgical size fractions. It was decided that the 2 series of sieves (Table 3.1) would provide reasonably narrow particle size fractions. Table 3.1 Chemical assays (%) of pyrite size fractions. Size Fraction, µm Fe Cu S Al 2O <0.09 <0.09 <0.09 <0.09 SiO < <0.42 <0.42 <0.4 <0.4 48

71 Table 3.2 Chemical Assays (wt %) of chalcopyrite size fractions*used in the tests. Size Fraction, µm All Sizes SiO CaO MgO Fe Cu S *see Appendix 3.1 for assay before the cleaning stage Quartz was ground under the same conditions as the model sulphide minerals, subsequently cleaned with hydrochloric acid and rinsed with tap water to natural ph. The quartz was sized and dried. The cleaning stage ensured all size fractions measured zero advancing contact angle. The cleaned quartz was used in chalcopyrite flotation tests in which a higher pulp density was required, and in contact angle measurements on mixed particle beds with chalcopyrite and quartz Sample Preparation (Natural ore) A porphyry copper ore supplied by Kennecott-Utah Coppertone Concentrator (KUCC), USA, was used in the ore test work. Crushed ore (< 2.4 mm) was blended and riffled into 2 kg samples. A grind calibration was carried out with 2 kg feed samples to establish the optimum grinding time to realise a d 80 of 220 µm (similar to plant operation). Three ore samples were ground for 5, 10 and 15 minutes and a size distribution analysis was carried out to determine the d 80 for each grinding time (Figure 3.2). An approximate grinding time of 12 minutes was estimated from the graph and a fourth sample was ground under the same conditions for twelve minutes. The size distribution after grinding, as well as the Cu and S assays per size fraction, are shown in Table 3.3. Bulk Mineral Analysis (BMA) by QEM Scan confirmed that the sulphur being minerals were chalcopyrite and pyrite only. The results of the ore test work are reported in Chapter 6. 49

72 d 80, µm Table 3.3 Size distribution for 2 kg of KUCC ore ground for 12 minutes (10 rods, 9.3 kg). Size Fraction, µm All Sizes Size Distribution, wt% Cu, wt% Typical recalculated Cu, S, wt% wt% Typical recalculated S, wt% Grind Time, mins Figure 3.2 Particle size for 80% passing as a function of grinding time for 2 kg of KUCC ore with 9.3 kg of stainless steel rods (10) Methods Contact Angle Manipulation Chalcopyrite-amyl Xanthate System In order to study the effect of contact angle on the flotation response of a given size fraction (Chapter 4), there was a need to vary the advancing contact angle of each size fraction in the feed. This was achieved by manipulating the contact angle of individual size fractions by oxidation at elevated or ambient temperatures (in the case of freshly prepared chalcopyrite) and by collector conditioning. Freshly prepared size fractions of chalcopyrite were hydrophobic and needed to be oxidised to decrease the advancing contact angle to lower values that could then be increased by collector addition. Selected size fractions were thermally oxidised at 150 o C to 50

73 achieve surface alterations. The duration at this temperature varied from 20 minutes to several hours depending on the degree of surface oxidation desired. Some chalcopyrite samples were kept in storage, which resulted in surface oxidation and a reduction in contact angle. The oxidised chalcopyrite size fractions were conditioned with collector as discussed further below. For each size fraction a mass of 50 g was conditioned with collector of concentration between 20 and 400 g/t (potassium amyl xanthate), at ph 10 for 1 hour. Some flotation feed samples were prepared by mixing size fractions into a bulk sample of 500 g and conditioning with collector at ph 10 for 1 hour, and then separating into size fractions by wet sieving. Each size fraction was then filtered, dried at ambient temperature, and the advancing contact angle measured on 10 g of sample. The remainder of the size fraction was then used in the flotation test, conducted immediately after measurement of the feed contact angle. It was difficult to increase the contact angle of the -20 µm size fraction above 70. High contact angles for this size fraction were achieved by conditioning the particles in acidic ph (1-2) for 1-2 hours with stirring, followed by filtration. The ph was adjusted using sulphuric acid. At least two bed volumes of water were used during filtration to remove residual acid. The filtered sample was immediately conditioned with collector solution at ph 10 for 1 hour at 10% solids with stirring. Alternatively, a fresh sample was ground in the presence of a collector and immediately floated (pre-flotation stage for chalcopyrite) and separated into size fractions. Some of the details of the experimental conditions to achieve the contact angle manipulation for each size fraction of chalcopyrite are included with test results in Chapter 4. The contact angle manipulation procedures were the same except where specifically indicated otherwise. It was not possible to target a specific contact angle by adding a certain amount of collector. The advancing contact angle on chalcopyrite that was achieved with a particular collector addition depended on the sample history, whether it had been freshly prepared, oxidised at elevated temperature or under ambient conditions. Samples oxidised at elevated temperatures had progressively lower capacity to adsorb collector as the duration of thermal oxidation increased. After the contact angle manipulation stage, the samples were 51

74 filtered and dried at ambient temperature, followed by advancing contact angle measurements as detailed further below Chalcopyrite-DTP System Selected size fractions were thermally oxidised at 150 o C to achieve surface alterations. Other selected samples were used directly from storage, the storage time being noted. After surface oxidation or storage, the advancing contact angle of each size fraction was increased by collector conditioning with sodium dicresyl dithiophosphate (DTP) (S8989). A mass of 50 g of each size fraction was conditioned with collector at ph 10 for 1 hour. The sample was then filtered, dried at ambient temperature, and the advancing contact angles measured on 10 g of each size fraction sample. The remainder of the size fraction was then used in the flotation test, conducted immediately after measurement of the feed contact angle. Details of the experimental conditions to achieve the contact angle manipulation for each size fraction of chalcopyrite are given in Table 3.4. These samples were tested in Chapter 5 and the test numbers correspond to those reported in that chapter. Table 3.4 Contact angle manipulation for chalcopyrite size fractions. Size Fraction, µm a b b a a a a a b Test 2 c c c c c c c c c 3 c c c c c c a a a 4 a a a a a a c c c a - storage (6 months), then conditioned with 2 g/t DTP at ph 10 for 1 hour, b - thermally oxidised (2 hours), then conditioned with 2 g/t DTP at ph 10 for 1 hour, c - storage (6 months), then conditioned with 20 g/t DTP at ph 10 for 1 hour Pyrite-amyl xanthate/dtp systems In the case of pyrite destined for flotation (results in Section 5.4), each size fraction was activated separately in the presence of 500 g/t CuSO 4.5H 2 O at ph 10, followed by the addition of collector to achieve different advancing contact angle values. A mass of 50 g of each size fraction was conditioned with collector at ph 10 for 1 hour. The sample was then filtered, dried at ambient temperature, and the advancing contact angle measured on 10 g of each size fraction sample. The remainder of the size fraction was then used in the flotation test, conducted immediately after measurement of the feed contact angle. Details of the experimental conditions to achieve the contact angle manipulation for each size fraction of pyrite are given in Table 3.5. Potassium amyl xanthate (KAX) and sodium 52

75 dicresyl dithiophosphate (DTP) collectors were used to increase the contact angles as indicated in Table 3.5. These samples were tested in Chapter 5 and the test numbers correspond to those reported in that chapter. Table 3.5 DTP. Contact angle manipulation for copper activated pyrite using KAX and Size Fraction, µm a a a a a a a a a Test 6 b b b b b b b b b 7 c c c c c c c c c a - conditioned with 50 g/t KAX at ph 10 for 1 hour, b - conditioned with 300 g/t KAX at ph 10 for 1 hour, c - conditioned with 30 g/t DTP at ph 10 for 1 hour Contact Angle Measurement Washburn Technique The method chosen for the measurement of contact angle on the powdered sample was the Washburn technique (Washburn, 1921). This method was chosen for its simplicity and the fact that several measurements can be carried out in a short time (Chau, 2009a) with good reproducibility (Muganda et al., 2008). The Washburn method makes use of capillary pressure to drive a liquid at an observable rate through a packed bed of particles in a capillary tube. The wetting velocity is then related to the advancing contact angle. Two approaches are commonly used, height versus time (height of the liquid front) and weight versus time (imbibition rate of measuring liquid). This method gives the advancing contact angle, only. The two approaches give the same advancing contact angle under appropriate conditions (Labajos-Broncano et al., 2001; Labajos-Broncano et al., 2003). When the measuring liquid front rises through the particle bed, the rate of penetration is related to the contact angle by dh dt LV 4 r cos h (3.1) where dh/dt is the penetration rate of the liquid into the pores with penetration depth, h, depending on the surface tension γ LV, liquid viscosity η, pore size (expressed by radius r) and advancing contact angle θ, and t is the time (Kviteck et al., 2002). To calibrate the pore 53

76 radius, r, which is usually unknown, a completely wetting liquid (such as cyclohexane with assumed contact angle of zero) is used and the contact angle is calculated as: (3.2) The weight versus time approach uses the rate of weight gain by a powdered sample suspended from an analytical microbalance. The method also uses two measuring liquids, one that is perfectly wetting (such as cyclohexane) for the determination of the material constant, and water, against which the contact angle is to be determined. Measurement with Milli-Q water utilises the following equation (Jackson et al., 2004): W 2 K 2 cos 2 t (3.3) where W is liquid weight gain, K is a material constant, and ρ is the density of the liquid. For the calibrating liquid which is fully wetting (cyclohexane), θ is assumed to be zero and equation (3.3) becomes W 2 K 2 2 t (3.4) From equation 3.4, K is calculated and substituted into equation 3.3 to calculate the advancing contact angle. The material constant, which is a geometric factor, reflects the porosity and capillary tortuosity, and is determined by particle size and bulk density (Jackson et al., 2004). The material constant K is determined by: (3.5) where r is the average capillary radius within the porous solid, and n is the number of capillaries in the sample (Rulison, 1996; Siebold et al., 1997). The determination of the values required to calculate K is difficult, and as a result they are grouped and calculated together by using a perfectly wetting liquid (Jackson et al., 2004). This method (Washburn) assumes (1) only laminar flow exists in the particle bed, (2) the effect of gravity is negligible, (3) the material packing structure is identical and remains constant during measurement. The weight/time approach was used in this thesis using the Wilhelmy Balance (Figure 3.3). 54

77 Figure 3.3 Wilhelmy Balance for contact angle measurements based on the Washburn technique. Method Description Two capillary tubes covered with a glass frit at one end were used for a single measurement (Figure 3.3). One capillary is required for each test liquid, water and cyclohexane. The glass frit was first covered with a filter paper to prevent particles from clogging the holes in the frit. An equal mass of particles from the same size fraction (approx. 2-3 g) was weighed and put into each capillary. The capillaries were electronically compacted for the same amount of time (about 3 minutes for the -20 µm fraction and about 2 minutes each for all other size fractions) and to the same bed height. Packing an equal mass of material to the same height in a capillary of the same diameter ensured that the packing density of the same size fraction in the two capillaries was the same. Proof of uniformity in packing density and bed porosity is provided in Chapter 4 by considering the volume of imbibed liquid for a given mass and size of particles. Contact Angle Measurement on Heterogeneous Mixtures of Particles The sensitivity of the Washburn method for the measurement of advancing contact angle on heterogeneous mixtures of hydrophobic and hydrophilic particles such as common in natural ore samples was tested using the µm size fraction of both chalcopyrite and clean quartz (advancing contact angle 0 ). The proportion of chalcopyrite conditioned with sodium dicresyl dithiophosphate (DTP) (with an advancing contact angle of 90 ) was varied from 0 to 100 % volume fraction of the mixture. The results of these measurements would provide a guide to the reliability of advancing contact angles measured directly on 55

78 concentrates and their potential use in the calculation of the contact angle of chalcopyrite in the concentrate of a natural ore. The results of this investigation are discussed in Chapter 6 (Section ) Advancing contact angles were measured on all combined concentrate size fractions from flotation tests with natural ore as discussed further below Equilibrium Capillary Pressure (ECP) Technique The wetting and dewetting characteristics of chalcopyrite and pyrite particles of the same size fraction and contact angle range were probed to establish whether there were differences in contact angle hysteresis. There was a need to probe the advancing and receding contact angles of chalcopyrite and pyrite at high and low contact angles for the µm size fraction. The method that can measure both the advancing and receding contact angles on particles is the Equilibrium Capillary Pressure technique. It is derived from the static method developed by Bartell (Dunstan and White, 1986), and uses a wetting liquid which rises in a column of the packed bed to a height H at which gravitational forces on the column of liquid balance the capillary pressure P L gh (3.6) An external pressure can be applied to oppose the capillary pressure and the pressure difference,, can be determined which just stops the capillary rise. Both the Washburn and ECP methods use the Laplace equation expressed as (Dunstan and White, 1986; Ralston and Newcombe, 1992; Diggins and Ralston, 1993) 2 LV r eff cos (3.7) where r eff is the effective capillary radius in the packed bed or porous material, which is unknown. It is standard practice to use a completely wetting calibrating liquid to determine the effective radius, which means using the same packed bed for this determination and this can lead to contamination. Using different packed beds introduces errors due to differences in packing density and size distribution of particles in the capillary. This method is the basis for the current equilibrium capillary pressure technique (ECP). When using two liquids, one perfectly wetting and the other non-wetting, the contact angle is given as: 56

79 cos (3.8) where ΔP is the Laplace pressure and γ is the surface tension of the measuring liquid (1) and perfectly wetting liquid (2). The technique can measure both advancing and receding contact angles on packed beds. Diggins and Ralston (1993) and Stevens (2005) improved the ECP technique to its current status. Method Description Pyrite was copper activated using 500 g/t Cu 2 SO 4. For each mineral (pyrite and chalcopyrite; µm size fraction), two samples (100 g each) were prepared, one with a high contact angle and the other with a lower value. This was achieved by addition of 200 g/t and 80 g/t of potassium amyl xanthate (KAX) at ph 10, respectively. After collector conditioning, the samples were filtered and left to dry in air overnight at ambient temperatures. Advancing and receding contact angle measurements were carried out using the ECP technique. Sufficient sample was put into the sample holder and the equipment set up as shown (Figure 3.4) and the measurements with milli-q water and cyclohexane as the calibrating liquid for each sample pair was carried out. The results are reported in Section A B C D E Figure 3.4 Assembled ECP apparatus ready for contact angle pressure measurement. The apparatus consists of (A) reservoir U-tube, (B) pressure transducer, (C) bleed valve, (D) reducing union and (E) the sample holder. (Stevens, 2005). 57

80 Flotation Tests Single Minerals: Chalcopyrite and Pyrite, 2% Solids After the measurement of contact angles on all size fractions, a flotation feed sample (100 g) was constituted (Table 3.6) and the sample was tested in flotation in a 5 dm 3 flotation cell (Figure 3.5), (Runge et al., 2003a). The cell impeller diameter was 10.5 cm and the flotation cell had dimensions of 20x20x13.5 cm for the length, width, and height, respectively. All flotation tests were carried out in this cell. The time interval from sample preparation to flotation did not exceed 20 hours, including repeat tests. Table 3.6 Size distribution of flotation feed (chalcopyrite and pyrite). Particle Size Fraction, µm % mass in size fraction, ±0.5% Typical* Recalculated feed, % Nominal Size Range Fine Intermediate Coarse Very coarse *calculated from flotation test products Each test was repeated twice, and the corresponding concentrates with time were combined for particle sizing. A third/fourth test was carried out if there were significant differences in the concentrate recoveries with time. Concentrates were taken at 1, 3, 5, and 8 minutes cumulative time intervals with a scraping rate of one every 10 seconds. The conditioning time at ph 10 was 1 minute and after addition of frother another 1 minute, giving a total of two minutes before air was introduced. Polypropylene glycol (PPG425) was used as frother. A 5% aqueous solution of the PPG425 was prepared. Analytical grade sodium hydroxide (1 molar) was used to adjust the ph. Figure 3.5 Bottom driven flotation cell with manual scraper for concentrate removal. 58

81 The following conditions were kept constant: a superficial gas velocity, J g of 0.3 cm/sec (Gas flow rate: 7.2 l/min), an agitation speed of 1200 rpm, a solids concentration of 2%, frother addition of 10 ppm at the beginning of the test and another addition of 10 ppm after collection of the second concentrate (necessitated by a decrease in froth stability after the second concentrate), and a froth depth of 1 cm. In order to probe the effect of Jg on the floatability components, some tests were carried out with a Jg of 0.5 cm/sec (Chapter 4, Section 4.6). All concentrates and tails were separated into size fractions to extract size-bysize recovery data Chalcopyrite Flotation at 30% Solids Chalcopyrite size fractions were conditioned individually with sodium dicresyl dithiophosphate (DTP) at ph 10 (Table 3.4, Tests 3 and 4). Contact angles of chalcopyrite size fractions were measured and a flotation feed sample of 100 g was constituted as before. The feed sample (100 g) with the same size distribution as in the single mineral test work (Table 3.6) was mixed with sized quartz (1900 g) to give an overall size distribution similar to the one for KUCC ore flotation feed (Table 3.3). Sodium hydroxide was used to increase the pulp ph to 10, instead of lime and no collector was added to the pulp because the chalcopyrite size fractions had already been manipulated to known contact angles prior to flotation. Each test was repeated three times to generate enough concentrate mass for sizing and assay. The results of these tests are reported in Chapter 5 (Section 5.3). Flotation tests were carried out under similar conditions to those of the KUCC ore outlined below Natural Ore Flotation As a standard procedure, a 2 kg ore sample was ground in a stainless steel rod mill with about 1.8 g of lime and 1 dm 3 of water for twelve minutes. The ground ore was immediately taken to a bottom driven flotation cell (Figure 3.5). The pulp density was adjusted by adding more water to give a solids content of 30%. The pulp was adjusted to ph 10 where necessary by adding more lime and subsequently conditioned with collector and frother. The collector used was S-8989 (sodium dicresyl dithiophosphate) and the frother was methyl isobutyl carbinol (MIBC). Four collector addition levels of 2, 15, 30, and 40 g/t dicresyl dithiophosphate (DTP) were used to produce discernible differences in hydrophobicity, while the frother addition was constant at 37.5 g/t or 60 µl in all tests. The 59

82 conditioning time for collector and frother was one minute each. The superficial gas velocity used was 0.3 cm/sec while concentrates were collected at cumulative times of 1, 3, 5, 8 minutes. Each test was repeated five times to generate mass for sizing, assay and contact angle measurements on the concentrates Chemical Assay and Mineralogical Analysis The feed, concentrates and tail samples were wet sieved into the size fractions shown in Table 3.3. The samples were wet-sieved to ensure that the surfaces of particles were cleaned of slimes. This was particularly important for concentrate samples destined for advancing contact angle measurements and surface analysis. The concentrate samples were filtered and dried in a dessicator under vacuum. Samples were collected for chemical assay. The chemical assays of the feed (Table 3.3) and flotation product size fractions allowed the recovery to be calculated as a function of particle size. The remainder of the concentrates for each size fraction were combined and used in direct contact angle measurements and surface analysis by ToF-SIMS (for the µm size fraction only) as explained further below. Mineralogical analysis of all feed size fractions and selected concentrate and tail samples of the ore was carried out by means of QEM-Scan. Three samples of the , , and +300 µm size fractions of the concentrate for the 15 g/t test were analysed to gain an understanding of the liberation characteristics, and the type of gangue in the concentrate. The liberation profile of the feed size fractions is shown in Figure 3.6, which showed that for size fractions up to µm, 80% of the chalcopyrite was at least in the 80% liberation class or greater. Liberation decreased above the µm size fraction, with the +300 µm fraction still at 50% of the chalcopyrite at least 80% liberation class or greater. This liberation profile has significance to the assignment of contact angle values for floatability components. 60

83 Cumulative Liberation Yield (%) Liberation by area, or apparent volume (%) Figure 3.6 Cumulative liberation yield as a function of percent liberation for chalcopyrite in KUCC ore feed size fractions. The bulk mineralogical composition of the ore feed is shown in Table 3.7. The only copper bearing mineral was chalcopyrite. There was a significant amount of iron-sulphide, relative to chalcopyrite, which complicated the determination of contact angle. The other gangue minerals were quartz, feldspars and mica. Table 3.7 Bulk mineralogical assay (Weight %) for KUCC ore by QEM Scan. Mineral Wt % Chalcopyrite 1.5 Pyrite 1.8 Quartz 30.2 Feldspars 36.7 Micas 24.4 Chlorite 1.1 Other sulphides 0.02 The presence of pyrite, in addition to chalcopyrite, minerals for which contact angle can change with collector addition complicated the direct measurement of contact angle, necessitating also the consideration of sulphur recovery representing both pyrite and chalcopyrite recovery. Since there were no other significant sulphur bearing minerals, sulphur recovery can be equated to the recovery of sulphide minerals, both chalcopyrite and pyrite. 61

84 Bubble size distribution measurement The bubble size was measured using the photographic method (Chen et al., 2001; Hernandez-Aguilar et al., 2005). A CCD camera was used to take images of bubbles in a viewing chamber as they rose from the pulp zone through a tube. Bubble size measurements with PPG425 as frother (used in single mineral work) were carried out with water, while measurements with MIBC were carried out with cleaned quartz, at 30 % solids, similar to the pulp density used in ore flotation tests. Images captured on the digital camera were processed using ImageJ ( The Sauter mean bubble diameter was then calculated using equation Data Analysis Recovery by Entrainment Recovery by entrainment in a single batch flotation test was estimated using the method of Ross, The rate of recovery by entrainment in the ith size fraction, E i (t) (g/min), at time t during flotation is (Ross, 1991): (3.9) where W(t) is the rate of recovery of water (g/min), and C mi (t) and C w (t) are the concentrates (g/l of pulp) of the particular mineral species and water at time t. X i (t) is a dimensionless transfer factor which is particle size dependent, and corresponds to the degree of entrainment (Lynch et al., 1981; Warren, 1985). Although X(t) is allowed to vary with time in the method by Ross (1991), to account for froth phase changes during the test, a constant value for X has been used in this work, which is only particle size dependent. To calculate the mass of particles recovered by entrainment in a test, the total flotation time was divided in intervals, corresponding to the time of collection of each concentrate. For each interval, the mass recovered by entrainment was calculated by means of Eq. 3.9, in which W(t) was experimentally measured, and the average concentrations of solids (C mi ) and water (C w ) in the cell were also calculated from the experimental recovery curves. Within each time interval, the mass of mineral particles recovered by entrainment was subtracted from the total recovery, obtaining in this way the recovery by true flotation. The dimensionless transfer factor X(t) was kept constant and approximated using the method by Savassi et al., (1998). The arithmetic mean of the minimum and maximum size 62

85 of a size fraction is taken as the particle size. A sample calculation for entrainment for the - 20 micron size fraction is shown below. Table 3.8 Sample calculation of recovery by entrainment and by true flotation for the - 20 µm size fraction Time, min Product Mass recovery, g Water Recovery, g Mass recovery by Entrainment, g (Value of X for -20 micron size fraction: 0.577) Mass recovery by true Flotation, g 1 Con Con *403.9*( /2)*/ = *134.6*( /2)/ = 0.13 ( ) = 6.78 ( ) = con *181.3*( /2)/ = 0.15 ( ) = con *288.6*( /2)/ = 0.21 ( ) = 0.29 Tail The method allows taking into account that the pulp in batch flotation is continuously depleted of particles due to both entrainment and true flotation. Recovery by entrainment was subtracted from the size by size recovery data to give recovery by true flotation, as shown in the last column of Table Recovery and Rate Constant The rate of particle recovery after entrainment was subtracted was calculated assuming a single floatable fraction and a non-floating fraction. The size by size recovery data, after entrainment was subtracted, was fitted to the first order rate equation: (3.10) using a least squares method where R(t) and R max are the recovery at time t and maximum recovery, respectively, and k is the distributed rate constant (rate constant of the floatable species). An example of the fitting process will serve to demonstrate the method. A sample of the +300 µm size fraction with a contact angle of 90 was floated. The experimental values are shown in Table 3.9. A fit to equation 3.10 gave corresponding cumulative model recoveries, with a constraint R max experimental recovery. The difference between the 63

86 experimental and model recoveries was squared and the sum of the square of differences was minimised. The fitted R max and rate constant are 73.5% and 1.30 min -1, respectively. Table 3.9 Experimental and fitted recoveries with time for +300 µm size fraction with 90 contact angle. Cumulative Flotation Time Experimental Cumulative Recovery Fitted Recovery Square of (mins) 0 (%) 0 (%) 0 Differences Sum of square of differences 9.4 Fitted R max= 73.5 R 2 = Fitted k= 1.30 k*= 0.96 A modified flotation rate constant, k*, which takes into account both maximum recovery, and rate constant was calculated (Agar et al., 1986; Sripriya et al., 2003). In this thesis, the modified flotation rate constant is called the undistributed rate constant, k*, and is calculated using: krmax k* 100 (3.11) Contact Angle and Surface Analysis Surface Analysis: X-ray Photo-electron Spectrometry (XPS) The variation of contact angle with particle size and its correlation to surface analysis using XPS was studied in order to understand the reason for the variation in advancing contact angle. Nine chalcopyrite size fractions were prepared without collector conditioning and each fraction was split into two. One sample was for contact angle measurement while the other for surface analysis. Surface analysis was carried out at the same time as contact angle measurement using the Washburn technique. Surface analysis was carried out using a Kratos Axis Ultra DLD XPS (monochromatic aluminium X-ray with an energy of ev, output power of 130 W, the binding energy scale from ev with a pass energy of 160 ev, and a step size of 500 mev and a dwell time of 55 ms). The samples were surveyed to identify all the elements on the surface and to determine the atomic concentration and chemical species. 64

87 Surface Analysis: Time of Flight Secondary Ion Mass Spectrometry (ToF- SIMS) An investigation was carried out to study the relationship between particle advancing contact angle and surface species of chalcopyrite conditioned with sodium dicresyl dithiophosphate. The study, which used the same chalcopyrite particles as in Chapter 5 of this thesis resulted in an empirical equation to correlate the advancing contact angle and surface species, which was independent of particle size. The methodology was trialled on the ore concentrate. The empirical relationship developed for the calculation of particle contact angle using ToF-SIMS intensities (I) for oxygen (O), sulphur (S) and collector (coll), in this case DTP, species on the surface of a particle is given by (Brito E Abreu et al., 2010) (3.12) This methodology was used to measure the contact angle of chalcopyrite particles in the concentrate of the µm size fraction for the 15 g/t flotation test. Only one size fraction was tested. This approach provided an indirect way to assess the accuracy of the directly measured contact angle of chalcopyrite particles in the concentrate using the Washburn technique, and thus inspire confidence in the reliability of the measured advancing contact angles of concentrate size fractions of the natural ore. A detailed description of the method appears elsewhere (Brito E Abreu, 2010). 65

88 CHAPTER FOUR 4. Correlating Particle Size and Contact Angle to the Flotation Response 1 Abstract The flotation response of chalcopyrite has been characterized as a function of particle size and advancing contact angle. The advancing contact angle of individual particle size fractions was controlled to different values. Flotation experiments were carried out under constant conditions of hydrodynamics and feed particle size distribution. A two-component model for each particle size fraction and contact angle was used to fit the experimental flotation data. The maximum recovery increased with advancing contact angle for each size fraction. It was found that the particle size fractions displayed flotation behaviour that was independent of each other. Particles within the same size fraction and within the same contact angle range displayed similar flotation behaviour, within experimental error, in different tests. Evidence for the apparent heterogeneity of the advancing contact angle within each particle size fraction was apparent in the appearance of a non-floating fraction. It was assumed that particles in the non-floating fraction had advancing contact angle values below the critical contact angle required for stable bubble-particle attachment. The flotation results were used to calculate the critical contact angle as a function of particle size, assuming a statistical distribution of contact angle values about a measured mean. Generally, the critical contact angle increases with particle size above 75 µm and for particles sizes below 20 µm. The recovery and rate constant data were collated into master curves of the undistributed rate constant and collection efficiency as a function of particle size for different contact angle ranges. These master curves, together with the critical contact angle, may be used to benchmark the flotation response of sulphide minerals and to infer an effective contact angle in a natural ore. 1 This chapter is to be published in modified form as the paper: Muganda et. al., (2010). Influence of particle size and contact angle on the flotation of chalcopyrite in a laboratory batch flotation cell. International Journal of Mineral Processing. 66

89 4.1. Introduction The influence of particle size and contact angle on flotation response was studied using model chalcopyrite particles floated under constant hydrodynamic conditions with a solids concentration of 2%. Model chalcopyrite particles were used in order to generate calibration data for use in benchmarking the flotation response of mineral particles in an ore. As with all calibration processes, accurate measurements are a prerequisite and the magnitude of error in the measurements needs to be quantified. The reliability of contact angle measurements on particles was critical to the success of this work. The experimental error in contact angle measurements, flotation recovery and rate constant is quantified. The characterisation of flotation response is dealt with in greater detail in this chapter, leading to the generation of calibration curves and the calculation of the critical contact angle of chalcopyrite. The calibration curves described in this chapter are validated in Chapters 5 and 6. The following hypotheses are tested in this chapter: Particle size fractions attain different contact angles when conditioned with collector in the same pulp. Particle flotation behaviour is related to the contact angle achieved. Hydrophobic mineral size fractions float independently of each other so that the first order kinetic model can be applied in the calculation of the flotation rate constant. This also allows the collation of flotation responses based on particle size fraction and contact angle ranges, which leads to the development of calibration curves that can be used to benchmark flotation response of minerals under the specific physical conditions. The maximum recovery of particles in a given size fraction and contact angle range can be modelled assuming a statistical distribution of individual particle contact angle around the mean. This leads to the determination of the critical contact angle for stable bubble-particle attachment. The flotation response of mineral particles of the same type, and within the same size fraction and contact angle range is the same irrespective of the hydrophobisation route, or type of collector used. 67

90 4.2. Error Analysis In this study, the magnitude of error involved in the advancing contact angle measurements needed to be evaluated to ascertain the sensitivity of the measurement method. Further, the error in the maximum recovery and rate constant for each size fraction was evaluated from results of repeated flotation tests, so that meaningful comparisons could be made between results from different tests. Error analysis is a key to the overall methodology of benchmarking flotation performance Error in Contact Angle The error in contact angle values was evaluated for each size fraction. Two samples of each size fraction were prepared, one with high and the other with low advancing contact angle (Table 4.1), by conditioning with 200 and 50 g/t of potassium amyl xanthate, respectively. This approach enabled probing of error for particles with different levels of hydrophobicity. Five repeated measurements using the Washburn technique on a homogenised sample of each size fraction were carried out. The average standard error for particles with high contact angle ( 78 ) was 0.7. Particles with low contact angle ( 55 ) displayed higher error values with an average standard error of 1.6. The magnitude of the error is small compared to that reported for the Equilibrium Capillary Pressure (ECP) technique (Stevens, 2005). Furthermore, the error is relatively higher in the fine and coarse particle size fractions, but not significant enough to suggest method bias. The measurement has adequate precision for the current study, with a maximum error of across all size fractions (Table 4.1). The error magnitude is in agreement with what has been reported in the literature for this method (Rulison, 1996; Stevens, 2005). Contact angle determinations within ±2.5 o were considered to be the same within experimental error. Table 4.1 Error in advancing contact angle measurement on chalcopyrite treated with potassium amyl xanthate (200 and 50 g/t, respectively for the high and low contact angle regimes, all size fractions conditioned together; Section ), n=5. Regime Size Fraction, µm High Contact Contact Angle, ( ) Angle Standard Error, ( ) Low Contact Contact Angle, ( ) Angle Standard Error, ( )

91 In order to gain better understanding of the method in terms of reliability of the contact angle values measured, further analysis was necessary. It is assumed that cyclohexane perfectly wets even highly hydrophobic particles. This assumption implies that the material constant is independent of particle hydrophobicity within the bed. The material constant, K, was plotted as a function of contact angle for selected particle size fractions (Figure 4.1a). The material constant appears to be independent of the measured contact angle of particles across all size fractions, in agreement with theoretical considerations (Rulison, 1996; Siebold et al., 1997; Jackson et al., 2004; Teipel and Mikonsaari, 2004). Figure 4.1 Material constant, K, of the chalcopyrite particle bed as a function of (a) contact angle for the -20 ( ), ( ), ( ) and +300 (o), µm size fractions, (b) particle size. The error in contact angle increases at the coarser end of the particle size distribution (Table 4.1). To investigate this phenomenon further, the material constant, K, was evaluated by averaging values from at least five advancing contact angle measurements on each size fraction, irrespective of the contact angle value. The material constant increases with size up to about 200 µm, above which it plateaus (Figure 4.1b). We observe that the volume of pores does not change with size, which means the change in K with size could be a result of a limiting capillary radius within the particle bed. This observation probably signals the approach to an upper particle size limit for the Washburn method, hence an increase in measurement error at the coarser end of the size distribution. Other researchers have used the Washburn method to measure advancing contact angles on particle size fractions up to 4 mm (Goebel et al., 2004). In this study the packing density was highly 69

92 reproducible as shown by the equal volumes of measuring liquid as discussed below; therefore an increase in the error may be attributed to the observed change in K with size. Repeatability of the particle bed packing density was confirmed by comparing the volume of measuring liquid imbibed into each capillary packed with the same mass and size fraction of particles to the same height. For each size fraction, the amount of liquid imbibed by the particles was calculated using the mass and density of the measuring liquids (Labajos-Broncano et al., 2003). The volumes of cyclohexane and water imbibed by the particle beds are approximately equal at both high and low contact angle values (Table 4.2). This result shows that the packing characteristics of the particle beds were repeatable in all samples. The volume of cyclohexane was consistently slightly lower than that of water due, probably, to its vaporisation ahead of the advancing liquid front (Table 4.2). The volume of liquid absorbed at low contact angles was the same irrespective of particle size. This suggests a constant pore volume irrespective of particle size for the same mass of particles. The total pore volume is a sum of the volume of individual capillaries in the bed. The capillary radius increases with particle size while the number of capillaries in the particle bed decreases. It appears the simultaneous variation of the capillary radii and number (equation 3.5) results in constant pore volume. The result may also indicate that the particle bed was packed close to the maximum packing density. However, when the contact angle increases, the amount of liquid imbibed is lower. This result suggests that hydrophobic particles may have attractive forces that result in a higher packing density, thus lowering the pore volume. Both measuring liquids gave the same volume, suggesting that hydrophobic interactions affect the packing characteristics of particles. The result may suggest that low contact angle (hydrophilic) particles are less attracted to each other and packing density is less. This may explain the observation that the standard error for low contact angle samples is higher. Table 4.2 Volume of measuring liquid imbibed by 2 g of chalcopyrite conditioned with potassium amyl xanthate (200 and 50 g/t, respectively for the high and low contact angle regimes, all size fractions conditioned together; Section ). Particle Size, µm Contact Angle, (±2.5 ) Vol. of Water, cm Vol. of cyclohexane, cm Contact Angle, (±2.5 ) Vol. of water, cm Vol. of cyclohexane, cm

93 Reproducibility of Flotation Tests The corresponding average recoveries displayed similar standard error trends as observed for the contact angle values (Table 4.3). The average standard error in R max is 1.5% for samples displaying high contact angles and 2.8% for samples displaying low contact angles. Table 4.3 Error in R max and k* for the same chalcopyrite size fractions with high and low contact angles as in Tables 4.1 and 4.2, n=3. All size fractions conditioned and floated together. Particle size, µm Contact Angle, (±2.5 ) R max, % Standard Error, %, k*, min Standard Error, min Contact Angle, (±2.5 ) R max, % Standard Error, % k*, min Standard Error, min The average standard error in k* is 0.1 min -1 for high contact angle regime and 0.04 min -1 for the low contact angle regime. The maximum standard error in k* obtained was 0.2 min -1. Undistributed rate constant determinations within ±0.2 min -1 were considered to be the same within experimental error. The results of the error analysis show that the methods chosen to measure the contact angles and flotation are sufficiently sensitive to give highly reproducible results Contact Angle Variation with Particle Size Contact angle on particles conditioned together in the same pulp varied with size fraction (Table 4.3). It was not immediately clear whether this variation was due to intrinsic effects of particle size on the contact angle measurement or due to genuine differences in surface species (Muganda et al., 2008). Significantly, the variation in contact angle values with particle size fraction was greater than the standard error in the contact angle measurement (Table 4.1). To explore reasons for this variation, a sample of chalcopyrite particles was prepared without collector addition. Each size fraction was analysed by XPS while at the same time contact angles were measured (Table 4.4). A multiple regression on the data (Table 4.5) shows that oxygen, copper, and sulphur are the principal species affecting the 71

94 advancing contact angle. Oxygen reduces the contact angle while copper and sulphur increase it. There is, however, co-linearity between the copper and sulphur species, a relationship that shows the two may exist together as copper sulphides. Sulphur has a more pronounced effect on the contact angle than copper. The analysis leaves us with just two important species defining the contact angle of the particle size fractions, i.e., O and S. The bulk of the oxygen exists as metal hydroxides and oxides, the abundance of which is particle size dependent, while the bulk of the sulphur was in the form of mono-, di- and polysulphides with little sulphate. These results confirm that sulphur species are mainly responsible for hydrophobicity in the collectorless flotation of chalcopyrite. Thus, while the contact angle varied with particle size fraction, the difference in contact angles with particle size is a result of genuine differences in the quantity of O and S bearing surface species. The result of this analysis settles the uncertainty that had arisen on the different contact angles measured on each size fraction and is important in affirming confidence in the measured contact angle as a measure of the degree of hydrophobicity. In this study, flotation recovery is correlated to advancing contact angle and it is necessary that the latter is measured accurately and truly reflects different surface species. Table 4.4 Contact angle and XPS atomic concentrations (%) (Appendix A2.2) on surfaces of chalcopyrite (in the absence of collector) size fractions. All particles conditioned together and then separated for contact angle and XPS measurement. Particle Size, µm Contact Angle, (±2.5 ) O C Cu Fe S Table 4.5 Regression coefficients for contact angle and XPS atomic concentrations. Variable CA O C Cu Fe S CA 1 O C Cu Fe S

95 4.4. Contact Angle, Particle Size and Flotation Response A series of flotation tests were carried out to probe the flotation behaviour of each particle size fraction as the contact angle varied. Attention was also given to the flotation response of particles hydrophobised differently but exhibiting similar advancing contact angle values within experimental error. In all flotation tests recovery by entrainment was subtracted from the total recovery. For flotation Test 1, a chalcopyrite flotation feed sample was prepared by conditioning the particle size fractions together with 300 g/t potassium amyl xanthate followed by contact angle measurement on the individual particle size fractions. A flotation feed sample was then constituted from the size fractions and floated under standard conditions (Section ). Detailed results of these tests are in Appendix A3.1. Particles less than 38 microns did not achieve high contact angle, probably because of the presence of oxidation products, as discussed previously. The fitted recovery profile with flotation time for each particle size fraction (Figure 4.2a) shows that a large portion of the floatable particles is recovered in the first minute. The recovery-time profile seems to justify only two components within each size fraction, a non-floating and a floating component. Although the contact angles of size fractions above 38 microns are high, there is still a non-floating fraction, due probably to the existence of particles with contact angles below the critical value. This conclusion is discussed further below. Table 4.6 Feed advancing contact angles and R max of manipulated chalcopyrite size fractions, Jg=0.3 cm/s. Particle Size, µm Test 1 Contact Angle, (±2.5 ) 74 b 69 b 90 b 90 b 90 b 90 b 90 b 90 b 90 b R max, ±3% Test 2 Contact Angle, (±2.5 ) 66 b 70 c 67 b 77 c 66 b 77 c 69 b 90 b 66 a R max, ±3% Test 3 Contact Angle, (±2.5 ) 36 a 64 b 72 c 79 b 61 a 52 a 62 c 71 c 72 c R max, ±3% a oxidised sample, no collector added; b collector conditioned sample; c freshly prepared sample with no collector. Experimental conditions are outlined in Sections and Flotation Tests 2 and 3 were carried out with feed samples constituted from size fractions individually manipulated to different levels (Table 4.6). A reduction in the nominal contact angle resulted in lower recovery (Figure 4.2b, c), and an increase in the non- 73

96 floating component (i.e., lower R max ). Noteworthy is the fact that particle size fractions displaying similar contact angles floated in a similar manner across the three tests. This is the case with the 20-38, 53-75, and µm size fractions. In the case of the (Test 1 & 2) and (Test 2 & 3) µm size fractions, the size fraction sample in one test was conditioned with collector while in the other it was conditioned in the absence of collector. These results demonstrate that when the advancing contact angle for a given particle size fraction is the same on the feed within ±2.5 o, the maximum recovery is also the same within experimental error (±3%), for chalcopyrite. The flotation behaviour of particles in different size fractions appears to be independent of each other in these tests in so far as changes in the contact angle of one size fraction do not change the flotation behaviour of the other particle size fractions. This important point is discussed further below. When the results of Tests 1-3 (Table 4.6) are considered together, it can be concluded that the recovery increases with advancing contact angle across all particle size fractions. The fitted curves in Figures 4.2a, b, and c show that it is reasonable to fit a single valued rate constant to describe the flotation rate of the floating component for this data. 74

97 Figure 4.2 Fitted cumulative recovery (%) with time for chalcopyrite size fractions (µm) in flotation tests. (a) Test 1, (b) Test 2, (c) Test ( ), (x), (+), ( ), ( ), ( ), ( ), (o), +300 ( ). 75

98 4.5. Flotation Independence of Mineral Size Fractions To further study possible interactions of particles in different size fractions during flotation, five further tests were carried out (Table 4.7). Some specific particle size fractions were intentionally and separately oxidised to reduce the contact angles as indicated by superscripts a or b against the contact angle values. The feed for each test was constituted by particles from size fractions that had been individually manipulated to alter the contact angles as indicated. Test 8 was carried out as a reference, with all size fractions oxidised to very low contact angles values. The recovery of each size fraction is closely related to the advancing contact angle. There was no recovery in Test 8 across all size fractions, showing that the measured contact angles were lower than the critical contact angle for stable bubble-particle attachment (Gontijo et al., 2007). The 53-75, , and +300 µm size fractions have similar recoveries in Tests 4 and 6, 5 and 6, 4 and 6, respectively (Figure 4.3). The undistributed rate constants (k*) are also in reasonable agreement, and are within the designated experimental error. For the and µm size fractions in Tests 4 and 6, 5 and 6, respectively, the particles were collector conditioned or collectorless, but show similar flotation behaviour, a result that suggests that particles of the same size fraction and exhibiting the same advancing contact angle float in the same way irrespective of whether it is collector or sulphur species which contributed to hydrophobicity, under the conditions of these tests. Apparently, particles float independently of each other as demonstrated by the fact that similar recoveries for each size fraction were obtained for similar contact angles in different tests. When the contact angle of a size fraction is changed, the floatability component (i.e., R max ) also changes. It appears floatability changes in one size fraction do not affect the floatability of another size fraction, at least within the particle size distribution and pulp density used in this study. For example, the µm size fraction in Tests 6 and 7 had similar contact angles (within experimental error) and the undistributed rate constant was the same, even though the contact angles of the other size fractions were different. The observation that particles float independently of each other allows the collation of flotation data from many different tests in which the contact angle for each size fraction is manipulated to different levels. This approach of analysing the flotation data is revisited further below. 76

99 Maximum Recovery, % Table 4.7 Advancing contact angles ( ) and undistributed rate constant k* (1/min) of chalcopyrite size fractions in different tests, Jg=0.3 cm/s. Particle Size, µm Test 4 Contact Angle, (±2.5 ) 36 a 90 b 90 b 67 b 40 a 79 c 90 b 90 b 37 a k*, min -1, ± Test 5 Contact Angle, (±2.5 ) 72 c 39 a 36 a 90 b 90 b 64 c 35 a 76 c 51 c k*, min -1, ± Test 6 Contact Angle, (±2.5 ) 54 b 64 a 79 c 69 c 80 c 63 a 78 c 72 c 39 a k*, min -1, ± Test 7 Contact Angle, (±2.5 ) 90 b 68 c 55 a 73 b 74 b 38 b 53 b 67 b 62 b k*, min -1, ± Test 8 Contact Angle, (±2.5 ) 30 a 10 a 0 a 12 a 5 a 16 a 0 a 3 a 11 a k*, min -1, ± a oxidised size fraction, no collector added; b collector conditioned sample after oxidation, c freshly prepared sample with no collector. (Preparation conditions are given in Sections and ) Test 4 Test 5 Test 6 Test Size Fraction, microns Figure 4.3 Maximum recovery as a function of particle size fraction. Size fractions manipulated to different contact angles shown in Table The Influence of Gas Flow Rate The maximum recovery of hydrophobic particles is achieved when the particles are given sufficient time to be recovered. In this study, the feed was assumed to have two components, floating and non-floating. The superficial gas velocity (J g ) used in tests to this stage was 0.3 cm/s. There was a need to assess whether there was any significant change in the floatability components when the superficial gas velocity was changed. Any change in the floatability components would require further tests to optimise the gas flow 77

100 rate and ensure that all floatable particles were recovered and accounted for in the maximum recovery. Flotation tests with an increased superficial gas velocity of 0.5 cm/s were carried out. Only results for particle size fractions with advancing contact angles within 5 of each other (i.e., within experimental error the same) across different tests are reported (Table 4.8). The results show that the maximum recovery for the two superficial gas rates is the same, within experimental error, for particles of a given size fraction exhibiting the same contact angle, also within experimental error. This demonstrates that the floatability components of a given size fraction are independent of the superficial gas rate, assuming there are two components in the feed. As expected, the rate constant increased with an increase in the superficial gas rate (Sripriya et al., 2003). This result further confirms that the bubble surface area flux is not limiting the recovery of the floatable particles under the standard test conditions. Table 4.8 Contact angle, flotation recovery and rate constants for selected size fractions. Effect of changing J g. Test A Test B Particle Size, µm J g, cm/s Contact Angle, (±2.5 ) Rmax, ±3% k*, min -1, ± Contact Angle, (±2.5 ) Rmax, ±3% k*, min -1, ± Contact Angle and Recovery The maximum error that was obtained in contact angle measurements was 2.4 (Table 4.1). The magnitude of error gives a contact angle range of about 5 within which we may expect to obtain the same flotation response for a given particle size fraction, within experimental error of flotation. This expectation seems reasonable when we consider the flotation results in Table 4.8. The R max values obtained with different superficial gas rates can be compared since the maximum recovery is the same in both situations. Across all the size fractions considered, the maximum recovery, within the experimental error of flotation (±3%), is the same for contact angle values within 5 of each other. Based on these observations, it was decided that the contact angle values be characterised in terms of ranges rather than specific values, so that the flotation recovery and rate constant of particles within a size fraction and contact angle range may be averaged across different 78

101 tests. This approach may be more discriminating in terms of the effect of contact angle on the flotation response. An analysis of the flotation tests data showed that the lowest contact angle range with non-zero recovery across all size fractions was The contact angle intervals for which flotation behaviour was characterised were established as 36-40, 41-45, 46-50, 51-55, 56-60, 61-65, 66-70, 71-75, 76-80, 81-85, and degrees. Several flotation tests were carried out to generate data for each contact angle range, and including the data displayed to this point. It should be pointed out that it was impossible to manipulate particle contact angles in such a way as to target a specific value. Eventually, sufficient flotation response data was generated for each contact angle range across all particle size fractions. For each size fraction, the cumulative recovery with time after subtraction of entrainment (Section ) within each contact angle interval from different tests was arithmetically averaged. Calculation of the maximum recovery, R max, the distributed rate constant, k, and the undistributed rate constant, k*, for the corresponding contact angle range was then carried out. The recovery increased with an increase in the contact angle as expected (Figures ). The flotation response of the - 20 µm size fraction remained suppressed (i.e., entrainment was the only recovery mechanism) up to a contact angle range of 61-65º (Figure 4.4). For contact angle values above 65º, the -20 µm size fraction recovery increased with contact angle. This result strongly suggests that fine particles have a higher critical contact angle to achieve floatability, a point which is discussed further below. The flotation response is particle size dependent and the degree of hydrophobicity required to achieve a given recovery increases with particle size above 20 µm. Contact angle values greater than 70 give nearly the same recovery-time curve for the µm size fraction (Figure 4.5). For this size fraction, an increase in contact angle above 70 does not yield much difference in terms of maximum recovery, but results in a greater rate of flotation (Figure 4.5). It appears the µm size fraction gives the best response to flotation. The recovery at the lowest contact angle varied with particle size. The flotation behaviour strongly suggests the existence of a critical contact angle for floatability. The theory of a critical contact angle is discussed further below in relation to the determination of the floatability components. 79

102 Figure 4.4 Cumulative recovery with time for different contact angle ranges for (a) -20, (b) 20-38, (c) µm chalcopyrite size fractions ( ), ( ), ( ), (+), ( ), (x), ( ). 80

103 Figure 4.5 Cumulative recovery with time for different contact angle ranges for (a) 53-75, (b) , (c) µm chalcopyrite size fractions ( ), ( ), ( ), (+), ( ), (x), ( ). 81

104 Figure 4.6 Cumulative recovery with time for different contact angle ranges for (a) , (b) , (c) 300+ µm chalcopyrite size fractions ( ), ( ), ( ), (+), ( ), (x), ( ). 82

105 4.8. Variation of the Rate Constant with Contact Angle for each Particle Size Fraction The flotation rate of the mineral particles increases with an increase in the advancing contact angle (Figure 4.7). The scale on the Y-axis is different for clarity of the graphs. The magnitude of the increase in rate constant with unit increase in contact angle is particle size dependent. A comparison of the rate constants dependence on contact angle has been discussed previously (Muganda et al., 2008). As the contact angle decreases, the distributed and undistributed rate constants diverge due to an increase in the proportion of non-floating component. It has been argued (Muganda et al., 2008) that the undistributed rate constant is a better indicator of the flotation response than the distributed rate constant. The recovery of particles is always less than 100% even when the measured contact angle on the feed size fraction is high (>80 ). The results presented bear this out. In order to fully characterise the flotation response in which there is a floating and a non-floating fraction, the undistributed rate constant is superior in that it includes the maximum recovery realised. Empirical equations relating the undistributed rate constant and the contact angle have been derived in such a way that contact angles can be calculated for a given rate constant (Table 4.9) or collection efficiency (Table 4.10). The collection efficiency, E coll, incorporates bubble sizes measured in Section and was calculated using equations 2.43, 2.45, and The regression coefficients are high, showing a good correlation between the advancing feed contact angle and undistributed rate constant. Table 4.9 Empirical equations for the calculation of contact angle from the undistributed rate constant, by size fraction. Contact Angle (±2.5 o ) = a(k*) b. Particle Size Fraction, µm a b R Table 4.10 Empirical Equations relating the contact angle and the collection efficiency, E coll. Contact angle (±2.5 o ) = a(e coll ) b. Particle Size, µm a b R

106 Figure 4.7 Variation of the distributed ( ) k, and undistributed (O) k*, rate constants with contact angle and particle size fraction. J g =0.3 cm/s, d b =0.48 mm, 1200 rpm, 2% solids. Fitted lines are calculated using equation and parameters shown in Table Calibration Plots for the Flotation Response The collation of all the flotation data presented in the previous sections helps to give a global picture of the flotation behaviour of particle size fractions within different contact angle ranges. The maximum recovery for a given contact angle range increases with 84

107 particle size up to the intermediate size fractions and decreases as the particle size continues to increase (Figure 4.8). The maximum recovery is highest at particle sizes between 20 and 105 microns depending on the contact angle range. The undistributed flotation rate constant varies with particle size for different contact angle ranges in a similar manner (Figure 4.9). There is considerable overlap of the error bars at the very coarse end of the particle size distribution for the undistributed flotation rate constant, and therefore the calibration is less discriminating for very coarse size fractions. The collection efficiency, E coll, as a function of particle size for different contact angle ranges (Figure 4.10) varies as the undistributed rate constant from which it is derived. The observed flotation behaviour of particle size fractions at different contact angle ranges can be explained in terms of the existence of a critical contact angle for stable bubbleparticle attachment. The shapes of these curves may suggest that the variation of the critical contact angle with particle size is a mirror image of the maximum recovery or undistributed rate constant curves. Where the maximum recovery and undistributed flotation rate constant are highest for any given contact angle range, the critical contact angle for the particular size fraction is probably lowest, giving rise to an optimum particle size for floatability. The critical contact angle is explored further below. The data constituting these master curves has been generated using single minerals that are fully liberated and floated under well defined conditions. It is possible that the flotation response obtained represents a standard or benchmark of flotation behaviour (calibration) of the same mineral in ores. 85

108 Undistributed rate constant, k*, 1/min Maximum Recovery, % Particle Size, microns Figure 4.8 Maximum recovery as a function of particle size for different contact angle ranges. Chalcopyrite single mineral. Agitation Speed=1200 rpm, Jg=0.3 cm/s Particle Size, µm Figure 4.9 Undistributed rate constant as a function of particle size for different contact angle ranges. Chalcopyrite single mineral. Agitation speed, 1200 rpm, Jg=0.3 cm/s. 86

109 Collection Efficiency, E deg deg deg deg deg deg deg Particle Size, microns Figure 4.10 Collection efficiency (E coll ) as a function of particle size for different contact angle ranges. Chalcopyrite single mineral. Agitation speed, 1200 rpm, Jg=0.3 cm/s Discussion The recovery of particles is dependent on the contact angle range and particle size fraction when the hydrodynamics are constant. For each particle size fraction and contact angle range, there is a non-floating component which decreases with an increase in the contact angle. Further, not all particles are recovered at the same rate even in situations where the bubble surface flux is not limiting. The existence of a non-floating component and the differences in flotation rates suggests contact angle heterogeneity within each particle size fraction of the feed as suggested by the work of Brito E Abreu et al., (2010). The maximum recovery varies with particle size fraction for any feed contact angle value. This suggests an intrinsic characteristic of particles that is size dependent, i.e., the critical contact angle. We may assume that the critical contact angle for floatability is constant for a given size fraction under constant physical conditions. If the critical contact angle is Z crit, the condition for particle recovery becomes: Particle contact angle Z crit All particles that satisfy the above condition have a probability of being recovered to the concentrate, while the particles with a contact angle less than the critical value cannot be recovered. The proportion of the floatable and non-floatable components in the feed 87

110 determines the maximum recovery. Within a size fraction/contact angle range of particles, there may be heterogeneity of both particle size and contact angle. The effect of heterogeneity in particle size can be reduced by working with very narrow particle size fractions, while heterogeneity in contact angle is more difficult to deal with. To demonstrate the effect of heterogeneity of contact angle, the following approach has been adopted. It was assumed that the contact angle of particles in a narrow size fraction are statistically distributed around the mean value of contact angle, which is the value measured by the Washburn technique. It has been assumed also, that, during flotation, particles in the size fraction with contact angles below the critical contact angle for floatability are non-floating, and therefore: (Non-Floating Fraction (Z i, Z criti )) = 100-R maxi [%] (4.1) where Z i and Z criti are the nominal and critical contact angles for the size fraction i, respectively. The non-floating fraction has been calculated for each size fraction/contact angle range of particles from the statistical distribution, as the cumulative distribution from 0 to Z crit (Figure 4.11). The gamma function is a two-parameter function with a mean and a mode. When the mean is equal to the mode, the gamma function approximates the normal distribution. As the mean and the mode of the distribution vary, there is skewness to the left or right, making this function suitable to describe contact angle distribution at both high and low contact angle regimes. It was assumed that the measured nominal contact angle using the Washburn technique is the mean, while the mode was varied to fit the data (Figure 4.11). This approach seems reasonable and physically meaningful when we consider contributions to the wetting characteristics of a packed bed. The contact angle measured on a particle bed is a surface area weighted mean of individual particle contact angle values. The shape of the contact angle distribution curve is assumed to vary with mean contact angle (Figure 4.11), but this variation is the same across all size fractions. The gamma distribution is given by (4.2) where α, λ, are the shape and inverse scale parameters, respectively. The mean is given by α*λ, while the mode is given by (α-1)λ. 88

111 Figure 4.11 Contact angle distribution in a size fraction of the feed ( µm, Z crit =55 ). Gamma distribution, measured contact angle = mean (Mean contact angle value indicated on the top left hand corner). The area to the left of Z crit (vertical broken line) represents the non-floating component in each contact angle range. The amount of the non-floating fraction decreases, and R max increases, as the mean contact angle increases. Contact angles in the distribution were constrained between 0 and 100, which are the minimum and maximum possible contact angles for the materials used (chalcopyrite-amyl xanthate system) (Sutherland and Wark, 1955). The mean of each contact angle range was taken as the nominal contact angle. A system of linear equations (equation 4.1) was written for each size fraction and nominal contact angle, in which the experimental R max 89

112 values were compared to the theoretical (model values, derived from the frequency distribution of contact angles). The following frequency distributions have been chosen for contact angles (Figure 4.11): If mean > mode,, and (4.3) If mean < mode,, and (4.4) A least squares method was used, allowing the critical contact angle Z crit and the distribution parameters α and λ to vary to minimise the fitting error. The experimental and model (from the fitting procedure) maximum recovery values had correlation R 2 value of 0.93 when compared across all size fractions and contact angle ranges (Figure 4.12). These results suggest that the contact angles of particles of any size fraction may be statistically distributed around the measured mean in the same way across all size fractions. What is not yet clear, however, is whether the same statistical distribution can be fitted to the contact angles of particles of similar sizes but of different minerals. It should also be noted that a normal distribution of contact angle values about the mean was also evaluated but this approach gave unrealistically high contact angle values (>120 o ) for high mean contact angle values. The issues of calibrated flotation behaviour, the critical contact angles and the influence of mineral specific gravity are explored further separately, in an attempt to test the applicability of the obtained parameters to different mineral collector systems (Chapter 5). The critical contact angle for stable bubble-particle attachment is a function of particle size (Table 4.11). The fine fraction has a high critical contact angle, which decreases for the intermediate size fractions, and increases with particle size above 75 µm. The trend is very similar to what has been reported previously (Crawford, 1986; Crawford and Ralston, 1988), with a minimum critical contact angle of about 48 for particles between 38 and 75 µm, increasing sharply as particle size increased above 150 µm or below 20 µm. It is considered that the principal problem in the flotation recovery of fine particles less than 20 m is due their high critical contact angle and the difficulty in furnishing contact angle values uniformly higher than the critical value. 90

113 Maximum Recovery (Experimental), % The unsized maximum experimental recovery was calculated for each contact angle range by taking into account the particle size distribution in the feed and the maximum recovery of each particle size fraction and contact angle range. This was compared with the unsized model recovery at the nominal contact angles, the correlation having an R 2 value of 0.98 (Figure 4.13). It may be possible to estimate an average contact angle for the unsized feed from the unsized maximum recovery for specific feed particle size distributions. Table 4.11 Calculated critical contact angle, Z crit, using the Gamma distribution model for contact angle values around the nominal (measured) value. Particle Size, µm Critical Contact Angle, ( ) When the distributed rate constant is calculated using the equation relating the contact angle and the undistributed rate constant for each size fraction (Table 4.9), the undistributed rate constant for particles exhibiting the critical contact angle value gives a value of 0.53±0.13 min -1 for all particle size fractions. The physical meaning, if any, of a uniform rate constant across the size fractions when particles have a contact angle equal to the critical value is not clear Figure 4.12 Maximum Recovery (Model), % Experimental maximum recovery versus Gamma distribution model maximum recovery. R 2 =

114 Unsized Maximum Recovery, (Experimental), % Unsized Maximum Recovery, (Model), % Figure 4.13 Experimental versus Model unsized maximum recovery at feed contact angles (mean) of 38 ( ), 53 ( ), 63 ( ), 68 ( ), 73 ( ), 78 (O), 88 ( ), R 2 = Conclusions The variation of contact angle with particle size for similarly treated chalcopyrite particles is a result of genuine differences in surface chemistry with particle size, with sulphur bearing species contributing positively to the contact angle and oxygen bearing species reducing the contact angle. Both the maximum recovery and rate constant increase with an increase in the advancing contact angle within any particle size fraction. The maximum recovery (i.e., floatability component) is not affected by the superficial gas velocity, but the rate constant increases with an increase in the superficial gas velocity. The flotation response varies with particle size as the contact angle is increased. Flotation rates increase rapidly with an increase in the contact angle for particles in intermediate size fractions, while only modest increases in rate constant are realised for very coarse size fractions. The flotation response of particles within the same size fraction is the same, 92

115 within experimental error (R max ±3%, k*±0.2 min -1 ), over a narrow contact angle range of about 5 under the conditions of these tests. Particles within different size fractions apparently float independently of each other in that changes in the contact angle of particular size fractions does not affect the floatability of other size fractions, under the conditions of these tests. The distributed and undistributed rate constants converge at high contact angles and diverge when the contact angle decreases due to an increase in the proportion of the nonfloating component. The undistributed rate constant is a more consistent and an accurate indicator of flotation performance based on particle size and contact angle. The maximum recovery, undistributed rate constant and collection efficiency of particles in different size fractions and in different contact angle ranges may be collated into curves which have the potential to be used for benchmarking the flotation response of minerals and to infer the contact angles of particles in the pulp based on flotation behaviour under well defined conditions. Mineral particles within the same size fraction achieve different contact angles even when they are subjected to the same chemical treatment. The contact angles of individual particles appear to be statistically distributed about a mean value as suggested by the work of Brito E Abreu et al., (2010). The apparent statistical distribution of contact angles gives rise to differences in flotation behaviour in terms of maximum recovery and rate constants within the same particle size fraction and contact angle range, as they are defined in this chapter. The critical contact angle for floatability of mineral particles varies with the particle size fraction. Fine particles less than 20 microns have a critical contact angle of ~71, while for particles above 75 microns, the critical contact angle increases with the particle size fraction. 93

116 CHAPTER FIVE 5. Effect of Collector Type, Pulp Density and Particle Specific Gravity on the Flotation Response 2 Abstract Chalcopyrite, conditioned with sodium dicresyl dithiophosphate (DTP), was floated under standard hydrodynamic conditions. The advancing contact angle values of the flotation feed and products were measured and the flotation response benchmarked against a calibration previously established for the chalcopyrite-amyl xanthate (KAX) system. Furthermore, the flotation response of pyrite, separately conditioned with potassium amyl xanthate and sodium dicresyl dithiophosphate (DTP), was also evaluated under the same standard hydrodynamic conditions. When the advancing contact angle of the chalcopyrite conditioned with DTP was similar (within 5 o ) to that of chalcopyrite conditioned with potassium amyl xanthate the flotation response was, within experimental error, the same. For both chalcopyrite and pyrite, heterogeneity of the advancing contact angle within the feed size fractions was demonstrated by significant differences in contact angle values of the flotation concentrate and tailings size fractions. For both chalcopyrite and pyrite, the recalculated feed contact angle was equal to, within experimental error, the measured feed contact angle demonstrating that the mean contact angle of the chalcopyrite and pyrite particles remained constant through both flotation and sample preparation under the test conditions. The flotation response of chalcopyrite, for both low and high advancing contact angle regimes, at 2% solids (w/w) was the same, within experimental error, as that at 30% solids (w/w) in the presence of silicate gangue, suggesting non-interaction of this gangue mineral with chalcopyrite under the test conditions. The operational advancing contact angles inferred for pyrite using the calibration established for the chalcopyrite-amyl xanthate system were, however, lower than the measured feed advancing contact angles, while the maximum recovery of pyrite was also lower than for chalcopyrite for the same feed advancing contact angle values, in the contact angle range less than 80. The difference in flotation response for the same feed contact angle values was interpreted in terms of a difference in critical contact angle value for stable bubble-particle attachment of these two mineral types. The critical advancing contact angle values of the pyrite size fractions, for both collector types, were higher than the critical advancing contact angle values of chalcopyrite for size fractions above 20 microns. This difference in critical advancing contact angle was attributed to the difference in mineral specific gravity between chalcopyrite and pyrite, rather than to differences in the distribution of contact angle values about the measured mean value in the feed. 2 This chapter is to be published in modified form as the paper: Muganda et. al., (2010). Inferring an operational contact angle on particles by flotation: Part 1 Single Minerals. International Journal of Mineral Processing. 94

117 5.1. Introduction The flotation characterisation of the chalcopyrite-xanthate system resulted in the development of calibration curves of the undistributed rate constant and collection efficiency versus particle size at different contact angle ranges (Figures 4.9 and 4.10 in Chapter 4). A modelling of the maximum recovery was carried out which resulted in the calculation of critical contact angles for stable bubble-particle contact for each size fraction. The applicability of these calibration curves is now evaluated against the chalcopyrite-dicresyl dithiophosphate (DTP) system at 2 and 30% (w/w) solids in the pulp, and the pyrite-xanthate and pyrite-dtp systems at 2% (w/w) solids. The results presented in this chapter are useful in strengthening the understanding of the flotation behaviour of particles. The hypotheses tested in this chapter are as follows: Particle flotation behaviour is related to the contact angle achieved. The maximum recovery of particles in a given size fraction and contact angle range may can be modelled assuming a statistical distribution of individual particle contact angle around the mean. This leads to the determination of the critical contact angle for stable bubble-particle attachment. There is conservation of floatability components through flotation as demonstrated by Washburn advancing contact angle measurements on flotation feed and products. When the conditions for first order kinetics are met with respect to hydrophobic particle concentration in the pulp, the influence of pulp density of a non-interacting gangue mineral on the flotation rate constant is insignificant, so that flotation behaviour at 2% solids is the same as that at 30% solids, the bulk being noninteracting gangue. The flotation response of mineral particles of the same type, and within the same size fraction and contact angle range is the same irrespective of the hydrophobisation route, or type of collector used. It is possible to infer an operational contact angle of hydrophobic mineral particles by benchmarking against the calibration, which can be confirmed by direct and indirect measurement of the contact angle. 95

118 5.2. Flotation Response of the Chalcopyrite-DTP System In this section, the flotation response of the chalcopyrite-dtp system (Section ; Table 3.4) is compared with the calibration curves developed previously (Chapter 4). For each size fraction the flotation response of high and low advancing contact angle regimes was probed, coupled with advancing contact angle measurements on the flotation products. The recovery - time profiles (Figure 5.1) for chalcopyrite size fractions conditioned with DTP to different contact angles (Table 5.1) shows that the maximum recovery is a function of both particle size and advancing contact angle. In the two tests, low and high contact angle regimes for each size fraction were probed. Apparently, the recovery plateaus at longer flotation times for almost all size fractions and feed contact angles such that a two component model of floatable and non-floatable particles is reasonably assumed. While only a single (mean) contact angle is measured in the feed there exist under most conditions at least two floatability components (i.e., R max significantly less than 100%). The non-floatable component, which reports to the tailing, has a significantly lower advancing contact angle than the floatable component in the concentrates, based on direct contact angle measurements of the products (Table 5.1). Implications of this finding are discussed further below. Table 5.1 Contact angle values of flotation products and the measured and recalculated advancing contact angles ( ) for DTP conditioned chalcopyrite before and after flotation, and the corresponding maximum recovery and undistributed rate constants. Low (Test 1) and high (Test 2) advancing contact angle regimes. (Table 3.4; Section ) Test 1 Test 2 Size Fraction, µm First concentrate, (±2.5 ) Second concentrate, (±2.5 ) Tail, (±2.5 ) Recalculated feed, (±2.5 ) Measured feed, (±2.5 ) R max, ±3% k*, ±0.2 min First concentrate, (±2.5 ) Second concentrate, (±2.5 ) Tail, (±2.5 ) Recalculated feed, (±2.5 ) Measured feed, (±2.5 ) R max, ±3% k*, ±0.2 min

119 Heterogeneity and Conservation of the Advancing Contact Angle The approach adopted in this study is to assume two floatability components in the feed for each size fraction, a floating and a non-floating component. The advancing contact angles of flotation products were measured to ascertain the basis of the separation and establish if there was conservation of advancing contact angle values. Results of the advancing contact angle measurements carried out on the original feed and the flotation products demonstrate that the advancing contact angle of particles in the feed remains largely unchanged during both flotation and sample preparation under the experimental conditions (Table 5.1). The recalculated feed advancing contact angle is a weighted average of the advancing contact angle of the flotation products using the respective mass fractions. Within the experimental error of the contact angle measurement of ± 2.5 o (Chapter 4), the recalculated and measured feed advancing contact angle values are the same for virtually all size fractions and for both low and high advancing contact angle regimes (Test 1 and Test 2 in Table 5.1). 97

120 Figure 5.1 Cumulative recovery with time for chalcopyrite size fractions conditioned with DTP and floated at 2% solids (a) Test 1 (b) Test 2. The continuous lines are for equation 3.10, after subtraction of recovery by entrainment. 98

121 Furthermore, the concentrates collected in the first minute (first concentrate) have consistently higher advancing contact angle values than the second concentrate. The tailing, which is solely the non-floating component (Figure 5.1), has lower advancing contact angle than both concentrates, showing that the separation of particles is contact angle dependent. This, of course, is expected given the basis of separation in flotation. The results of the advancing contact angle measurements on the flotation products further confirm that individual particles within individual size fractions of the feed achieve different advancing contact angles when conditioned with collector under the same conditions. The heterogeneity of the advancing contact angle is probably due to differences in the abundance of active sites on the mineral surface (Piantadosi et al., 2000), at least. Differences in contact angle of particles separated by flotation have been demonstrated for the first time in this study and it is anticipated that future studies relating particle size, flotation rate constants and contact angle may lead to the determination of contact angle distribution within a given size fraction depending on flotation behaviour. This may represent a leap forward in the development of diagnostic tools for process control in mineral processing Benchmarking Flotation Response The undistributed rate constant for each size fraction in Tests 1 and 2 (Table 5.1) is clearly dependent on the advancing contact angle measured on the feed. To benchmark the flotation response of the chalcopyrite-dtp system against the chalcopyrite-amyl xanthate system, empirical equations relating the advancing contact angle to the undistributed rate constant determined for the chalcopyrite-amyl xanthate system (Chapter 4) were used to calculate an inferred operational advancing contact angle value. The measured feed and inferred operational advancing contact angles, based on the undistributed rate constant, are in good agreement for both the low and high advancing contact angle regimes, with R 2 values of 0.94 and 0.95, respectively (Figure 5.2). The result further confirms that the flotation behaviour of DTP conditioned chalcopyrite is the same as amyl xanthate conditioned chalcopyrite, provided that they have the same advancing contact angle (within 5 of each other). Liberated chalcopyrite particles exhibiting the same advancing contact angle, within experimental error measured using the Washburn method, but with apparently different surface functional groups exhibit the same 99

122 flotation behaviour, within experimental error. This result implies that the efficiency of the sub-processes of attachment and detachment for chalcopyrite are the same when the advancing contact angle falls within the same range (±2.5 o ). Figure 5.2 Inferred operational advancing contact angle compared with the measured feed advancing contact angle of DTP conditioned chalcopyrite size fractions with (a) low, R 2 =0.94 (Test 1) (b) high advancing contact angle (Test 2) regimes, R 2 =0.95. The critical advancing contact angle values by size fraction determined for the chalcopyrite-amyl xanthate system (Chapter 4) are now compared with the measured advancing contact angles of the flotation products of the chalcopyrite-dtp system (Figure 5.3). The advancing contact angle values of the concentrates are above the critical advancing contact angle for both high and low advancing contact angle regimes, while the tailing advancing contact angle values are at, or below, the critical value. The tailing advancing contact angle value in the intermediate (20-53 µm) and very coarse (+300 µm) 100

123 particle size fractions for the low advancing contact angle regime (Figure 5.3a) have significantly lower advancing contact angles than the critical value, probably due to the fact that these size fractions were intentionally thermally oxidised (Table 3.4) to achieve low contact angles in the feed (Table 5.1). For the high advancing contact angle regime (20 g/t DTP in Test 2) the advancing contact angles of the tailing size fractions are, within experimental error, the same as the critical advancing contact angle values determined for the chalcopyrite-amyl xanthate system. The higher collector addition in Test 2 is likely to give a narrower distribution of advancing contact angle (Brito E Abreu et al., 2009), so that the particles that do not float are at the critical value required for stable bubble-particle attachment. There is remarkable agreement between the chalcopyrite-amyl xanthate and chalcopyrite-dtp system, with flotation clearly separating on the basis of advancing contact angle, with particles with advancing contact angle lower than the critical value reporting to the tailing stream as the non-floatable component. It appears that the critical advancing contact angles calculated using the chalcopyrite-amyl xanthate system are valid for the chalcopyrite-dtp system, possibly implying that the critical advancing contact angle may be independent of the hydrophobising species at the molecular scale. The critical advancing contact angle is a function of the particle size, and possibly mineral specific gravity, as discussed further below. It could be concluded that the critical advancing contact angle and the flotation response may be independent of the hydrophobisation route for a specific mineral. Given a measured feed mean advancing contact angle and assuming a distribution of advancing contact angle values about the mean, then it may be possible to predict the maximum recovery (i.e., floatable components) of the mineral. 101

124 Figure 5.3 Advancing contact angles of concentrates and tails for DTP conditioned chalcopyrite with (a) low (2 g/t DTP, Test 1), (b) high (20 g/t DTP, Test 2) advancing contact angles compared with the critical advancing contact angle curve (chalcopyriteamyl xanthate system) Tests at Higher Solids Percent Flotation Response at High Pulp Density Flotation of ores is normally carried out at higher pulp density than that used in the study of the chalcopyrite-amyl xanthate (Chapter 4) and chalcopyrite-dtp single mineral 102

125 systems discussed thus far. This disparity of pulp density (i.e., % solids w/w) warranted further examination. The ultimate aim is to provide a useful benchmark against which ore flotation behaviour can be compared. The effect of pulp density on the flotation response was probed using chalcopyrite particles preconditioned with DTP (Table 5.2, Tests 3 and 4) for which the individual contact angle values were measured, and then mixed with clean, hydrophilic quartz. Two tests were carried out so that the flotation behaviour of individual size fractions at both low and high advancing contact angle regime was probed (Table 5.2, Figure 5.4). As in previous tests, there is clearly a floating and non-floating component in each size fraction. The higher the contact angle measured on the feed, the lower the non-floating component (100-R max ). The feed advancing contact angle values measured on each of the chalcopyrite size fractions, before mixing with quartz, in the two tests are shown in Table 5.2. Table 5.2 Flotation response of chalcopyrite with quartz at 30% solids (w/w). Chalcopyrite conditioned with DTP (2 and 20 g/t) prior to flotation. Size fractions conditioned individually, and floated together. (Section ; Table 3.4) Particle Size Fraction, µm Feed Advancing contact Test angle, (±2.5 ) 3 R max, ±3% E coll, ± Test 4 Feed Advancing contact angle, (±2.5 ) R max, ±3% E coll, ± Benchmarking against the Calibration Changes in pulp density may affect bubble size (Grau, 2006b), and therefore the bubbleparticle collision efficiency. The level of turbulent energy dissipation is also lower at higher solids percent, and both the bubble-particle attachment and detachment processes may be affected (Pyke et al., 2003) to an extent. 103

126 Figure 5.4 Cumulative recovery with time for chalcopyrite size fractions conditioned with DTP, floated at 30% solids (w/w), (a) Test 3 (b) Test 4. The continuous lines are for equation 3.10, after subtraction of recovery by entrainment. The Sauter mean bubble diameter was determined to be 0.57 mm at high (30% w/w) solids percent, and was constant through flotation. At low (2 % w/w) solids concentration, the Sauter mean bubble diameter was 0.48 mm, and was also constant through flotation. This difference in bubble size was incorporated into the calibration by considering the collection efficiency (E coll ), rather than simply the undistributed rate constant (k*). To benchmark the flotation response of the chalcopyrite at high solids percent, empirical equations relating 104

127 the advancing contact angle to the collection efficiency (based on the undistributed rate constant, Table 4.10) at 2% solids were used to calculate an inferred operational advancing contact angle in the feed. An inferred operational advancing contact angle was then calculated for each size fraction in Tests 3 and 4 (Table 5.2). The results show that, within the experimental error of flotation, there is very good agreement between the measured feed and inferred operational advancing contact angles (Figure 5.5), with an R 2 value of 0.98 for both Tests 3 and 4. Changes in energy dissipation (lower turbulence) at higher solids percent may benefit the stability efficiency, while also reducing the collision frequency due to reduced energy dissipation and increased bubble size. The net effect of these subtle and opposing influences seems to be minimal in any case. The conclusion that the percent solids (up to 30% w/w) do not affect the calibration may be valid in the specific case of a chemically inert and non-interacting gangue component such as quartz at ph10. Chalcopyrite and pyrite particles may not interact with quartz in the circumstances because both are negatively charged under alkaline conditions and changes in pulp viscosity may be minimal over this range. The conclusion is also tested in Chapter 6 in which a natural ore with chalcopyrite as the main copper mineral and poly-component gangue mineralogy is tested by flotation. 105

128 Figure 5.5 Operational advancing contact angle compared with measured feed advancing contact angle for chalcopyrite feed size fractions floated at 30% solids (w/w), (a) Test 3, (b) Test 4. R 2 =0.98 in both tests. 106

129 5.4. Flotation of Pyrite The calibration curves developed previously on the chalcopyrite-amyl xanthate system (Chapter 4) were used to benchmark the flotation response of pyrite under the same hydrodynamic conditions. The pyrite-amyl xanthate and pyrite-dtp systems were investigated to probe the effect of collector species and mineral type (e.g., specific gravity) on flotation response. For each size fraction, high and low advancing contact angle regimes with copper activated pyrite conditioned with potassium amyl xanthate (KAX) were probed (Table 5.3, Tests 5 and 6), while the flotation response of pyrite conditioned with dicresyl dithiophosphate (DTP) was tested at a low advancing contact angle regime only (Table 5.3, Test 7). Table 5.3 Advancing contact angles (CA) of pyrite flotation feed and products (First, second concentrates, and tail), and corresponding maximum recovery and undistributed rate constants. Corresponding R max values for chalcopyrite size fractions within the same advancing contact angle range as in Tests 5-7. (Section ; Table 3.5) Test 5 (KAX) Test 6 (KAX) Test 7 (DTP) Particle Size, µm First Concentrate CA, (±2.5 ) Second Concentrate CA,±2.5 ) Tail CA, (±2.5 ) Recalculated Feed CA, (±2.5 ) Measured Feed CA, (±2.5 ) R max, %, ± k*, min -1, ± First Concentrate CA, (±2.5 ) Second Concentrate CA, ±2.5 ) Tail CA, (±2.5 ) Recalculated Feed CA, (±2.5 ) Measured Feed CA, (±2.5 ) R max, %, ± k*, min -1, ± First Concentrate CA(±2.5 ) Second Concentrate CA,(±2.5 ) Tail CA, (±2.5 ) Recalculated Feed CA, (±2.5 ) Measured Feed CA, (±2.5 ) R max, %, ± k*, min -1, ± R max for chalcopyrite size fractions with the same corresponding contact angle range as in Tests 5-7, respectively Feed Contact Angle Range, ( ) R max, ±3% Feed Contact Angle Range, ( ) R max, ±3% Feed Contact Angle Range, ( ) < R max, ±3%

130 Apparently, only low advancing contact angle values could be obtained with the pyrite- DTP system, which would be a result of differences in the adsorption mechanism between copper sulphide and pyrite when conditioned with DTP (Fuerstenau et al., 1971). It is important to note firstly, that the recalculated feed advancing contact angle values in Table 5.3, based on the weighted measurements on the flotation products, is the same, within experimental error (±2.5 o ), of the advancing contact angle measured on the feed. This finding further supports the conclusion that the advancing contact angle of particles present in the feed remains largely unchanged through both flotation and sample preparation under the experimental conditions. The advancing contact angle values of the flotation products were measured, and, as observed in the case of the chalcopyrite-dtp system, the first concentrate was more hydrophobic than the second concentrate. The tailing, which is dominated by the nonfloatable component, had lower advancing contact angles than either concentrate. Differences in advancing contact angles of the flotation products demonstrate conclusively that there is heterogeneity of advancing contact angles within each size fraction of the feed, as observed in the case of the chalcopyrite-dtp system. Advancing contact angle heterogeneity may therefore be a characteristic of all mineral ensembles such that when an advancing contact angle is measured on a bed of particles, it is actually the surface area weighted mean of a statistical distribution. There appears to be a difference in the recovery time profiles for pyrite as compared to chalcopyrite. The recovery - time curves plateau at shorter flotation times for chalcopyrite (Figure 5.1) than pyrite (Figure 5.6). This phenomenon is more apparent when the pyrite feed contact angle is low (Fig. 5.6c). This suggests that the rate of pyrite flotation is lower than chalcopyrite for the same feed advancing contact angle, an issue that is discussed further below. Comparison was also made between the maximum recovery of chalcopyrite and pyrite for each size fraction exhibiting the same contact angle. The maximum recovery for pyrite of a given size fraction and advancing contact angle is less than that for chalcopyrite, for advancing contact angle values less than 80 (Table 5.3, Figure 5.7). 108

131 Figure 5.6 Cumulative recovery with time for pyrite size fractions (a) KAX, Test 5 - low contact angle regime, (b) KAX, Test 6 - high contact angle regime, (c) DTP, Test 7. The continuous lines are for equation 3.10, after subtraction of recovery by entrainment. 109

132 Considering that the flotation tests were carried out under the same hydrodynamic conditions, possible reasons for the apparent discrepancy between chalcopyrite and pyrite are (i) the pyrite size fractions have a wider distribution of advancing contact angle values about the mean, such that a greater proportion of pyrite particles are below the critical contact angle value required for stable bubble-particle attachment, and/or (Wierink et al.) the pyrite size fractions have higher critical advancing contact angle values than chalcopyrite. This may explain why pyrite has lower recovery than chalcopyrite for similar advancing contact angles within a specific size fraction. These possibilities are discussed further below. Figure 5.7 Comparison of maximum recoveries (%) for chalcopyrite and pyrite for each size fraction. The values are for advancing feed contact angles for pyrite and chalcopyrite in Table 5.2. The inference of an operational advancing contact angle based on the undistributed rate constant may provide further insight into the flotation behaviour of pyrite compared to chalcopyrite. Empirical equations (Table 4.9) derived from the chalcopyrite-amyl xanthate system were used to calculate an operational advancing contact angle for pyrite from the undistributed rate constant values. Generally, the inferred operational advancing contact angles are lower than the measured feed advancing contact angles (Figure 5.8) in all tests. 110

133 Figure 5.8 Inferred operational advancing contact angle compared with measured feed advancing contact angle for (a) pyrite-xanthate system with low advancing contact angle Test 5, (b) pyrite-xanthate system with high advancing contact angle Test 6, (c) pyrite-dtp system with low advancing contact angle Test

134 Lower inferred operational advancing contact angle implies that the calibration curves generated with the chalcopyrite-amyl xanthate system may not hold for pyrite due to mineral specific factors such as specific gravity, shape and roughness, etc, or due to a wider distribution of contact angle values about the mean value. The roughness of the mineral surface affects its wetting characteristics (Sedev et al., 2004; Krasowska and Malysa, 2007). In this study the advancing contact angle was generally used to characterize hydrophobicity. However, for successful bubble-particle attachment, the receding contact angle may play a significant role in determining the rate of intervening liquid film thinning, rupturing and receding. This particular issue is tackled in the following section Wetting and Dewetting Behaviour of Chalcopyrite and Pyrite It was decided to compare the wetting and dewetting behaviour of chalcopyrite and pyrite size fractions exhibiting the same advancing contact angle, within experimental error, of the Washburn technique (±2.5 o ). The high and low advancing contact angle regimes were probed for contact angle hysteresis of each mineral, obtained by subtracting the receding from the advancing contact angle (See Section for details). There does not appear to be significant difference in contact angle hysteresis for both minerals, suggesting that the wetting and dewetting behaviour is the same, based on advancing contact angle data for the micron size fraction (Table 5.4). This result appears to rule out differences in flotation behaviour between the two minerals due to differences in receding contact angle. Further, this exercise provided opportunity for comparison of the advancing contact angle values measured by the two common methods for powdered particles. There is general agreement in the advancing contact angles measured, given the relatively large error in contact angle measurement for the ECP method. 112

135 Table 5.4 Advancing and receding contact angles of feed chalcopyrite and pyrite conditioned with KAX, µm size fractions. (Section ) θ adv, ECP θ rec, ECP Washburn θ adv, ±2.5 o Chalcopyrite 65±8 61± ±6 75±7 82 Pyrite 58±6 55± ±3 83±5 87 θ adv, θ rec are the advancing and receding contact angles, respectively. Copper activated pyrite conditioned with DTP did not achieve high advancing contact angles. The maximum advancing contact angles achieved were in the medium to low range (50-60 o ). The measured feed and inferred operational advancing contact angles show scatter, especially in the intermediate size range (Figure 5.8c). Further discussion on the benchmarking of pyrite flotation response against chalcopyrite follows below Discussion From the literature, individual particle advancing contact angles in a size fraction may be statistically distributed around a mean. In Chapter 4, a gamma (Γ) statistical distribution of individual particle advancing contact angles around the measured mean was assumed following the work of Brito E Abreu et. al. (2010). Brito E Abreu et al. (2010) used the same particles as those used in the current study of the chalcopyrite-dtp system. The model developed for chalcopyrite can predict the maximum recovery of any size fraction when the feed advancing contact angle is known, assuming a particular statistical distribution of advancing contact angle values (Figure 4.11). The statistical distribution of advancing contact angles around the mean value may be particle size dependent, and/or mineral specific. Two alternative assumptions, in the consideration of pyrite, are considered in this section: (1) Pyrite size fractions have the same critical advancing contact angles as those calculated for chalcopyrite, and the distribution of individual particle advancing contact angle around the mean is the same irrespective of particle size fraction, (2) Pyrite size fractions have different critical advancing contact angle values than chalcopyrite size fractions, but the same statistical distribution (i.e., the same relationship between the measured mean value and the mode of the contact angle distribution in Figure 5.9) can be used to fit the flotation data for both minerals. Pyrite flotation behaviour was compared to that of chalcopyrite assuming a gamma statistical distribution of contact angle within a size fraction, and fitting the pyrite flotation 113

136 Mode Contact Angle, ( ) data using the same parameters determined previously (i.e., assumption 1). For the feed mean advancing contact angles shown in Table 5.3, a mode was determined using Figure 5.9, based on relationships derived from chalcopyrite-amyl xanthate system (Chapter 4). The procedure for the model fit has been dealt with in detail in Chapter Mean Contact Angle, ( ) Figure 5.9 Variation of the mean and mode advancing contact angle values for a sample, assuming gamma statistical distribution of individual particle contact angles. The mean is the value measured by the Washburn technique. The mean and mode contact angle values determine the contact angle distribution in a feed sample. When assumption 1 was applied, the model fit to the pyrite flotation data overestimated the maximum recovery for the mean contact angles measured using the Washburn method (Figure 5.10a), with an R 2 value of only This finding may suggest that either pyrite has a wider distribution of advancing contact angles around the mean value than chalcopyrite (i.e., Figure 5.9 does not hold for pyrite), and/or, pyrite has a higher critical contact angle possibly due to specific gravity differences between the two minerals. The latter possibility seems more likely given that chalcopyrite and pyrite particles exhibit similar contact angle hysteresis for the same mean advancing contact angle (Table 5.4), suggesting the same statistical distribution of contact angle values around the mean. 114

137 The possibility of similar advancing contact angle distributions in the feed for chalcopyrite and pyrite size fractions was evaluated by comparing the cumulative masses of tail, second and first concentrates (Table 5.1 and 5.3), in that order, and the corresponding cumulative mean contact angle for specific size fractions of the two minerals exhibiting the same mean contact angle in the feed, within the experimental error of the advancing contact angle measurement (±2.5 o ). There is close agreement between the recalculated cumulative mass - cumulative mean advancing contact angle relationships for the two different minerals (Figure 5.11), providing demonstration that the same statistical function may be applied to pyrite flotation data as for the chalcopyrite (i.e., Figure 5.9 applies for both minerals). Fitting to the pyrite flotation data under assumption 2 was attempted by allowing the critical contact angle to vary to obtain the best fit. A least squares method that minimised the difference between the model and experimental R max values was used. An example of the fitting procedure for the and +300 µm size fractions is shown in Table 5.5. The model was successful in predicting pyrite recovery (Figure 5.10b) with an R 2 value of 0.97 across all size fractions. These results appear to support the assumption that the advancing contact angles are statistically distributed around the measured mean contact angle in the feed in the same way for both pyrite and chalcopyrite, and that the pyrite size fractions generally have higher critical advancing contact angles than chalcopyrite. Thus, the lower recovery of pyrite compared with chalcopyrite, for the same feed advancing contact angle (Figure 5.7), in the range less than 80, may be explained by the higher critical contact angle for stable bubble-particle attachment in the case of pyrite. Pyrite has a higher specific gravity and therefore lower stability efficiency than chalcopyrite. Evidently, this difference in critical advancing contact angle is not driven by differences in surface functional groups but by differences in mineral specific gravity. 115

138 Figure 5.10 Experimental versus model maximum recovery for pyrite flotation, 2% solids. (a) Assumption 1, critical advancing contact angle for pyrite is the same as for chalcopyrite, R 2 =0.84, (b) Assumption 2, pyrite has a different critical advancing contact angle than chalcopyrite, R 2 =

139 Figure 5.11 Cumulative mass recovered into the tail, second, and third concentrate (total = recalculated feed) against the cumulative mean contact angle for pyrite (O) and chalcopyrite ( ) for (a) -20 (Tests 2 and 6), (b) (Tests 2 and 6), (c) (Tests 1 and 6), (d) +300 (Tests 2 and 6) µm size fractions from Tables 5.1 and 5.3. Tests compared with similar mean contact angles values in the feed. Table 5.5 Pyrite flotation data model fitting to a gamma statistical distribution with mean (measured) and mode contact angle. The critical contact angle, fixed for each size fraction, was allowed to vary to obtain the best fit. Size Fraction, µm Mean Contact Angle, ( ) Mode Contact Angle, ( ) R max (Model), % (Exptl, %) (Model-Exptl) R max ( R max) Σ( R max) 2 Critical Contact Angle, ( )

140 A higher advancing contact angle is required to recover pyrite particles in flotation at the same rate as chalcopyrite particles within the same size fraction. The critical advancing contact angles for pyrite and chalcopyrite particles have been determined in this work under the same specific hydrodynamic conditions (Figure 5.12). It should be noted that there is a significant difference in the critical contact angle between chalcopyrite and pyrite for all size fractions greater than the -20 micron size fraction. This supports the hypothesis that the difference in critical contact angle between chalcopyrite and pyrite is probably due to the difference in specific gravity between the minerals. For the -20 micron fraction, effects due to differences in specific gravity would be minimal, and there is only a small difference in the critical contact angle. Fine particles less than 20 microns, for both chalcopyrite and pyrite, have a higher critical contact angle for stable bubble-particle attachment than intermediate size fractions due probably to their lower kinetic energy (Scheludko et al., 1976). The reason fine particles generally have lower maximum recovery than intermediate sized particles is due to their higher critical contact angle and the difficulty in furnishing mean contact angle values well above the critical value. As the particle size increases above approximately 53 microns, the critical contact angle again increases for both chalcopyrite and pyrite due to decreases in the stability efficiency of the bubble-particle aggregate (Schulze, 1977). Changes in the hydrodynamic conditions may give different critical contact angle values. If these results reflect the general behaviour of mineral particles, then it should be possible to determine the maximum recovery of any fully liberated mineral in a specific size fraction when the feed mean contact angle and critical advancing contact angle values are known, and assuming the same statistical distribution of contact angle values as for the chalcopyrite-amyl xanthate system. The methodology needs to be tested against a natural ore, a difficult task tackled in the next chapter. 118

141 Critical Contact Angle, º Pyrite Chalcopyrite Particle Size, µm Figure 5.12 Variation of the critical advancing contact angle with particle size for pyrite and chalcopyrite Conclusions Chalcopyrite particles exhibiting the same mean advancing contact angle in the feed (within experimental error), but with different functional groups on their surfaces (i.e., amyl xanthate and dicresyl dithiophosphate) are recovered with a similar undistributed rate constant (within experimental error). This finding further validates the calibration developed using the chalcopyrite-amyl xanthate system, and extends its application to a different collector on chalcopyrite. The conservation of the advancing contact angle through flotation and sample preparation for a given size fraction, for both chalcopyrite and pyrite, was demonstrated by the very close agreement between the measured and recalculated feed contact angle values under the conditions of this study. The heterogeneity of contact angles in each feed size fraction, for both chalcopyrite and pyrite, was confirmed by measurements of contact angle on the flotation products. The concentrate contact angles values were always higher than the tailings contact angle values which, in turn, were always at or lower than the critical contact angle which is specific for each mineral. The critical contact angle required for 119

142 chalcopyrite for stable bubble-particle attachment is apparently independent of surface species on the mineral surfaces (i.e., amyl xanthate and dicresyl dithiophosphate). Under the conditions of the present study, the flotation response of chalcopyrite particles floated at pulp densities of 2 and 30% solids, with quartz as non-interacting gangue, was the same within experimental error. This finding may allow benchmarking of chalcopyrite flotation behaviour from an ore against the calibration developed for the chalcopyrite-amyl xanthate system. The undistributed rate constants of pyrite were lower than for chalcopyrite exhibiting the same feed advancing contact angle, within experimental error, while the maximum recovery of pyrite was also lower for the same feed advancing contact angle, in the contact angle range less than 80. This difference was attributed to differences in the critical advancing contact angle for pyrite and chalcopyrite due to differences in mineral specific gravity, rather than to differences in surface groups, or contact angle distribution about the measured mean value. It may be possible to predict the maximum recovery of a specific liberated mineral in a specific size fraction with knowledge of the feed mean advancing contact angle and critical contact angle values, and assuming the same statistical distribution of contact angle values as for the chalcopyrite-amyl xanthate system. The same statistical function may be applied to both pyrite and chalcopyrite flotation data to determine the distribution of contact angle values given a mean contact angle measured in the feed. The methodology developed may apply to all minerals. It may be possible to predict the maximum recovery of a liberated mineral as a function of size fraction when the feed contact angle is known, or, conversely, to infer an operational contact angle based on particle size, flotation rate constant, and critical contact angle of a mineral under the specific test conditions used in this study 120

143 CHAPTER SIX 6. Benchmarking a Natural Ore Against Calibration Curves 3 Abstract A porphyry copper ore containing chalcopyrite as the principal copper bearing mineral, and pyrite as the only other sulphide mineral, was treated in batch flotation tests under well defined physical conditions and at three different collector additions. The size-by-size flotation response was benchmarked against established calibration curves to infer an operational contact angle of the sulphide minerals (i.e., chalcopyrite alone, and chalcopyrite and pyrite combined) as a function of particle size. The inferred operational contact angle values of the sulphide minerals (i.e., chalcopyrite and pyrite combined) were validated by independent measurements of contact angle on the concentrates and, in the case of chalcopyrite, by an indirect approach using Time of Flight Secondary Ion Mass Spectrometry (ToF-SIMS). Recovery, flotation rate, and inferred operational contact angle for both chalcopyrite and sulphide minerals increased with collector addition across all size fractions, with the intermediate and coarse size fractions benefitting the most from increased collector addition. The directly measured and inferred operational contact angles were in reasonable agreement, with an R 2 value of 0.7 across all size fractions. There was good agreement between the advancing contact angle values determined using ToF-SIMS and those calculated from direct contact angle measurement on the µm size fraction for chalcopyrite. A method for the determination of particle contact angle using flotation response has been developed, and is expected to lead to better flotation process diagnostics and modeling. 3 This chapter is to be published in modified form as the paper: Muganda et. al., (2010). Inferring an operational contact angle on particles by flotation: Part 2 Natural Ore. Minerals Engineering. 121

144 6.1. Introduction The characterisation of the flotation response of the chalcopyrite-xanthate system has led to the development of calibration curves of the undistributed rate constant, k*, and collection efficiency, E coll, against particle size at different contact angle ranges (Chapter 4). In Chapter 5, it was demonstrated that the flotation response of the chalcopyrite-dtp system benchmarked against the calibration curves was the same for similar advancing contact angles. Further, the flotation behaviour of chalcopyrite particles conditioned with DTP and floated at a typical pulp density of 30% solids (chalcopyrite quartz mixture) was the same as at 2% solids (100% chalcopyrite). The result suggests that the same flotation response may be expected from a natural chalcopyrite ore, apart from the presence of composite particles, which were not considered in the single mineral work. The inferred operational contact angles of pyrite-xanthate and pyrite-dtp systems were lower than the feed advancing contact angle across all size fractions, suggesting that the specific gravity of a mineral aggregate influences the flotation rate constant. The flotation rate constant of mineral particles that exist as composites may reflect the influence of the average specific gravity of all the minerals in the composite, and not necessarily that of the value mineral. Issues of the degree of liberation and its effect on composite particle specific gravity may introduce significant error, especially for the coarse size fractions. The flotation of chalcopyrite at 30% solids served as a preamble to benchmarking a natural ore against the calibration curves, which is the objective of this chapter. Three tests with collector additions of 2, 15, 30 g/t sodium dicresyl dithiophosphate were carried out. These collector additions, which are significantly different, afforded opportunity to probe the contact angle achieved by various particle sizes as the amount of collector was increased Natural Ore Flotation and Inference of an Operational Contact Angle Chalcopyrite Flotation Response The flotation behaviour of chalcopyrite obtained in laboratory tests (Section ) on the natural ore is typical of plant practice in terms of the recovery dependence on particle size (Figure 6.1) and flotation time (Figure 6.2). The final (unsized) recovery increased with collector addition, reaching 71, 80, 85, and 87 % for 2, 15, 30, and 40 g/t of DTP, respectively. The test at 40 g/t DTP produced the same results, within experimental error, 122

145 Maximum Recovery, % as the 30 g/t test, showing that the collector was in excess at this high addition. The plant generally uses approximately 20 g/t of DTP for this ore (Triffett et al., 2008). It is evident (Figure 6.1) that coarse particles benefit more from increased collector addition, a conclusion similar to that observed in other mineral systems (Polat and Chander, 2000; Vianna, 2004). The increase in recovery with collector addition may be explained in terms of an increase in surface coverage by collector and contact angle of the coarse particles. Fine and intermediate particles apparently respond in flotation at lower collector additions (Trahar, 1981b) g/t 15 g/t 30 g/t 40 g/t Particle Size, µm Figure 6.1 Maximum recovery (R max ) of chalcopyrite as a function of particle size at different collector additions for KUCC ore. Collector (DTP) added at 2 ( ), 15 ( ), 30 ( ), and 40 g/t (o). Frother 37.5 g/t MIBC, 1200 rpm, 30% solids (w/w). Increased collector addition, and hence surface coverage of collector, is explained in this work as more particles being hydrophobised to contact angles above the critical value for stable bubble-particle attachment. Apparently, the flotation behaviour of particle size fractions below 53 microns (Figure 6.1) was nearly the same in the four tests. The fine to intermediate size fractions appear to preferentially adsorb collector, possibly because they have a higher surface area to solution volume ratio than the coarse particles. Another less likely possibility is that the degree of liberation is also higher for the fine to intermediate 123

146 size fractions, thus more surface is available to adsorb collector. As the collector addition is increased, coarse particles adsorb more collector and become more hydrophobic, hence there is an increase in recovery with an increase in collector addition. The maximum recovery was highest for the micron size fraction in all three tests. This is the optimum size fraction for flotation under the test conditions. 124

147 Figure 6.2 Cumulative Cu recovery with time, natural copper ore, (a) 2 g/t, (b) 15 g/t, (c) 30 g/t dicresyl dithiophosphate collector. For size fractions:( ) -20, ( ) 20-38, ( ) 38-53, (+) 53-75,( ) , ( ) , ( ) , (o) , ( ) +300 The continuous lines are for equation Recovery by entrainment has been subtracted. 125

148 The decrease in recovery with an increase in size above 53 microns may be attributed to the increase in the critical contact angle with particle size (Table 4.11). A smaller number of particles have a contact angle above the critical value when collector addition is low, hence recovery is lower. The degree of liberation of the particles may also affect the contact angle achieved by a mineral aggregate, especially when the mineral particle exists as a composite, in association with hydrophilic gangue. As the particle size increases, the stability of the bubble particle aggregate in agitated pulps decreases due to an increase in mass, thus a higher contact angle is required for bubble-particle aggregate stability. A fit of the first order rate equation to the experimental recovery with time data after subtraction of entrainment was very good across all size fractions (Figure 6.2). It appears that the recovery with time reaches an asymptotic value (R max ) across all size fractions, giving justification to the two floatability components model (i.e., a floating and a nonfloating fraction). The bulk of the floatable component is generally recovered in the first minute of flotation (Figure 6.2), even for the -20 µm fraction Benchmarking Chalcopyrite Flotation Behaviour The flotation behaviour of the chalcopyrite in the natural ore closely resembles that observed in the single mineral experiments (Chapters 4 and 5), in terms of the recoverytime profile (Figure 6.2). The objective of this study is to infer the contact angle of the hydrophobised chalcopyrite particles in the ore. The flotation behaviour of chalcopyrite is benchmarked against calibration curves previously developed for the chalcopyrite-amyl xanthate system (Chapter 4). Further, the difference in pulp density between the single mineral work and ore work necessitated taking into account inevitable changes in bubble size distribution. In the single mineral work with 2% solids, the Sauter mean bubble diameter was 0.48 mm while that with 30% solids (chalcopyrite mixed with clean quartz) was 0.57 mm (Chapter 5). The Sauter bubble size measured was also 0.57 mm throughout flotation in the current ore test work. The collection efficiency, E coll, was used to benchmark the flotation behaviour of chalcopyrite in an ore. This conversion from rate constant to collection efficiency (with the use of equations 2.43, 2.45 and 2.46) eliminates the bubble surface area flux as a variable arising from differences in bubble size distribution (determined in Section 3.2.4). Empirical equations (Table 4.10) relating the contact angle to the collection efficiency, based on the undistributed rate constant, were used, so that for a given value of the 126

149 Collection Efficiency, Ecoll collection efficiency in the ore work, an inferred operational contact angle value can be calculated. Table 6.1 Collection efficiency, E coll, and inferred operational contact angles of chalcopyrite in KUCC ore feed after different collector additions. Particle Size Fraction, µm g/t E coll, ± Inferred Contact Angle, ±2.5 > g/t E coll, ± Inferred Contact Angle, ±2.5 > g/t E coll, ( ) ± Inferred Contact Angle, ±2.5 ( ) > The concept of the benchmarking procedure is illustrated in Figure 6.3, where the collection efficiencies are plotted against the calibration curves to infer an operational contact angle. This illustration, however, does not give a single contact angle value as obtained with the empirical equations, but a narrow range of contact angles, and the situation can be difficult if there is overlap of error bars as in the coarse end of the calibration g/t 15 g/t 30 g/t Particle Size, microns Figure 6.3 Collection efficiencies of particle size fractions for 2, 15, 30 g/t ore flotation tests benchmarked against calibration curves of the collection efficiency, E coll, against particle size for different contact angle ranges (from Chapter 4, Figure 4.10). 127

150 An increase in the collector addition above 2 g/t did not increase the collection efficiency significantly for particle size fractions less than 53 microns. Generally, the collection efficiency increased with collector addition for size fractions above 53 microns only, although the increase was small for the very coarse particles. As the particle size increases, the collection efficiency increases and reaches a maximum for the micron size fraction, and decreases with an increase above this size fraction. The maximum collection efficiency is obtained with the micron size fraction at the highest collector addition. The inferred operational contact angles increase with an increase in collector addition for each size fraction above 53 microns. It appears that particles less than 53 microns achieved high contact angles even at low collector additions. Higher collector addition did not result in higher contact angles. The natural hydrophobicity of chalcopyrite (corroborated by high contact angles for collector-less single mineral chalcopyrite in Table 4.4) may have contributed to the high, inferred operational contact angles, even at starvation levels of collector addition. This is more so, considering that, the particles in these size fractions, are fully liberated and therefore, have a larger surface area available for collector adsorption. Assuming that these results are a true indication of the response of particles to conditioning with collector, the collection efficiency of copper bearing particles increases with an increase in the degree of hydrophobicity of individual particles within the same size fraction. It is important to note that the inferred operational contact angle is the contact angle on the feed of the valuable mineral, as both the ore and the benchmark (single mineral chalcopyrite-amyl xanthate system) collection efficiencies are calculated from the undistributed rate constant, taking into account both the floating and non-floating components in the feed. The flotation behaviour of chalcopyrite in a natural ore may be affected by other minerals in the ore such as pyrite. In the following section, the flotation behaviour of sulphides (i.e., chalcopyrite and pyrite) is considered Sulphide Flotation Response The natural ore contains a considerable amount of pyrite (see the bulk mineralogical analysis, Table 3.7). Pyrite may be activated by copper and becomes hydrophobic in the presence of a collector. The presence of pyrite in the ore complicates an already difficult task in terms of determining the hydrophobicity of chalcopyrite in the ore. It is difficult 128

151 enough to deal with one floatable species due to existence of heterogeneity of both contact angle and floatability components, but when there are two different minerals floating at the same time, the task becomes onerous. It was decided, therefore, to consider sulphur recovery, representing chalcopyrite plus pyrite recovery, referred to below as sulphide recovery. QEM-Scan analysis justified this assignment, as the only significant sulphur bearing minerals are chalcopyrite and pyrite. The maximum recovery of sulphides as a function of particle size at 2, 15, and 30 g/t collector addition is shown in Figure 6.4. At 2 g/t collector addition, chalcopyrite recovery (Figure 6.1) is higher than sulphide recovery across all size fractions. Although the recovery of chalcopyrite for the fine to intermediate size fractions is high at this collector addition, it appears that pyrite flotation remains suppressed possibly because there is insufficient collector to hydrophobise it to a contact angle above the critical value. At higher collector additions, chalcopyrite and sulphide recovery are the same, within experimental error, up to 53 µm. Above 53 µm, the sulphide recovery is higher than chalcopyrite recovery, which may be explained by increased recovery of pyrite as it becomes sufficiently hydrophobic to float at higher collector additions. These results seem to suggest that chalcopyrite preferentially adsorbs collector to pyrite, or that collector adsorption may be the same, but that pyrite requires more collector to attain sufficient hydrophobicity to float, since, as discussed in Chapter 5, pyrite has a higher critical contact angle for stable bubble-particle attachment. The former is probably the most likely explanation. The recovery-time profiles for sulphides appear similar to chalcopyrite (Figure 6.2), but at higher collector additions (15 and 30 g/t DTP), the flotation rate of the intermediate to coarse size fractions of the combined sulphides appears higher (Figure 6.5). When the assays of the concentrates and the calculated weight percent and volume fractions of the chalcopyrite, pyrite and non-sulphide gangue minerals are considered (Table 6.2), it becomes evident that pyrite recovery remained low at 2 g/t collector addition and was higher across all size fractions greater than 20 µm for higher collector additions. A multiple regression on the three components of the concentrates (chalcopyrite, pyrite, nonsulphide gangue) shows that the amount of chalcopyrite is negatively correlated to the amount of non-sulphide gangue up to about 210 µm, while pyrite becomes significant in its association with gangue minerals above that size. The recovery of sulphides is significantly higher than that of chalcopyrite alone at the coarse end of the size distribution 129

152 Rmax, % for higher collector additions, and this is borne out by the calculated concentrate compositions shown in Table Benchmarking Sulphide Flotation Response The flotation response of the sulphides (chalcopyrite and pyrite) was benchmarked against the calibration previously developed for the chalcopyrite amyl xanthate system using the collection efficiency as before (Table 6.3). The inferred operational contact angles generally increase with collector addition and decrease with particle size above 38 µm. Hydrophobic particles in the size fractions between 20 and 300 µm appear to have the same collection efficiency (and thus flotation rate) for the 15 g/t collector addition: the same flotation behaviour is noticed for the 30 g/t collector addition for the same range of sizes. Also significant is the difference between the flotation rate of chalcopyrite and sulphides for the coarse size fractions at 30 g/t collector addition. It appears that pyrite is hydrophobised to high contact angles and its recovery is higher than that of chalcopyrite. Composite particles of pyrite and other non-sulphide gangue minerals like silicates appear in abundance as shown by the concentrate composition (Table 6.2) g/t 15 g/t 30 g/t Particle Size, µm Figure 6.4 Maximum recovery (R max ) of sulphide minerals (chalcopyrite and pyrite) as a function of particle size at different collector additions 2 ( ), 15 ( ), and 30 ( ) g/t DTP for KUCC ore. 130

153 A comparison of the inferred operational contact angles of chalcopyrite and sulphide minerals is made in Figure 6.6. For the test at 2 g/t collector addition, chalcopyrite generally has higher contact angle values than the combined sulphides (Figure 6.6a). For the test at 15 g/t, there appears to be good agreement between the inferred contact angles, showing that more pyrite was hydrophobised to higher contact angles, thus floating at the same rate as the chalcopyrite particles (Figure 6.6b). The sulphides (combined) attain contact angles equal to, or greater than that of chalcopyrite, at the 30 g/t collector addition (Figure 6.6c), a result that implies more pyrite is hydrophobised to higher contact angle values than chalcopyrite. 131

154 Figure 6.5 Cumulative sulphur recovery with time, natural copper ore, (a) 2 g/t, (b) 15 g/t, (c) 30 g/t dicresyl dithiophosphate collector. For size fractions:( ) -20, ( ) 20-38, ( ) 38-53, (+) 53-75,( ) , ( ) , ( ) , (o) , ( ) +300 The continuous lines are for equation Recovery by entrainment has been subtracted. 132

155 Table 6.2 Cu and S Assays of KUCC ore concentrates from flotation tests with calculated weight % and volume fractions of chalcopyrite (chp), pyrite (Pyke et al.) and non-sulphide gangue (NSG). Size Fraction, µm Cu, wt % S, wt% wt% chp g/t Vol. chp wt% py Vol. py g/t 30 g/t wt% NSG vol. NSG Cu S wt% chp Vol. chp wt% py Vol. py wt% NSG vol. NSG Cu S wt% chp Vol. chp wt% py Vol. py wt% NSG vol. NSG Table 6.3 Collection efficiency, E coll, and inferred operational contact angles (CA) of sulphides (chalcopyrite and pyrite) in KUCC ore feed after different collector additions. Particle Size Fraction, µm g/t E coll, ± Inferred CA, ±2.5 > g/t E coll, ± Inferred CA, ±2.5 >90 > g/t E coll, ± Inferred CA, ±2.5 >90 > An attempt has been made to validate some of the contact angles inferred using the benchmarking procedure from calibration curves developed for the chalcopyrite-amyl xanthate system (Chapter 4). The inferred operational contact angles are based on the feed contact angle for each size fraction before particle separation by flotation. The approaches for validation included an indirect method using surface analysis by ToF-SIMS specifically addressing chalcopyrite, and direct contact angle measurements on the ore feed, concentrate, and tails addressing the sulphide minerals. These approaches are discussed further below. 133

156 Figure 6.6 A comparison of the inferred operational contact angle of chalcopyrite and sulphide (chalcopyrite and pyrite) for different size fractions at collector additions of (a) 2 g/t, (b) 15 g/t, (c) 30 g/t DTP. 134

157 6.3. Validation of Inferred Operational Contact Angles Surface Analysis Technique for Contact Angle Measurement A contact angle/species intensity (ToF-SIMS) correlation was developed for the chalcopyrite-dtp system using the same chalcopyrite sample used in Chapter 5. The developed methodology was trialled on the µm size fraction (Section ) of the concentrate for the test at 15 g/t on the ore in the current study. Only mineral particles of chalcopyrite were considered in the analysis. The Washburn advancing contact angle measurement method incorporates the effects of particle surface roughness and shape, so that the correlation between the advancing contact angle and ToF-SIMS species intensity may be valid. The predicted average contact angle of chalcopyrite particles in the concentrate is 74º over 33 different chalcopyrite particles (Brito E Abreu, 2010). The standard error of prediction is 14º. The distribution of contact angles of this group of particles is displayed in Figure 6.7. The median is 69º and the mode is around 70º. The standard deviation is 21. The Washburn technique can only measure advancing contact angles up to 90. Some chalcopyrite particles in the ore appear to have adsorbed more collector than that observed in the calibration of the advancing contact angle and ToF-SIMS intensity correlation using single mineral chalcopyrite, resulting in contact angle values greater than 90. All particle contact angles above 90 degrees are in the extrapolation range and it is not yet clear how accurate these values are. As discussed previously, it is unlikely that the chalcopyrite particle advancing contact angle would be above 100 for the mineral-collector system. Further, the ToF-SIMS correlation holds for oxygen, sulphur and collector species present on chalcopyrite in the ore system being the same as on the single mineral system from which it is derived. If for example, oxygen is derived from silicates from which the oxygen yield is different from oxygen bearing species on chalcopyrite (e.g., iron oxide) in the calibration stage, discrepancies may occur. The measured average contact angle by ToF-SIMS appears reasonable when compared with the inferred contact angle of 77 (Table 6.1, µm, 15 g/t), but the inferred contact angle is for the feed, not the concentrate only. This value is compared with calculated value from direct advancing contact angle measurements further below. 135

158 Figure 6.7 Distribution of the advancing contact angles measured using ToF-SIMS of chalcopyrite particles in the concentrate of the µm size fraction of the test at 15 g/t. The contact angle calculated using surface analysis is for the floatable component only, while the tail component of the valuable mineral could not be determined due to very low concentration of the chalcopyrite particles that did not float. Another approach involving direct contact angle measurements on concentrates from all flotation tests was considered, and is discussed further below Advancing Contact Angle Measurements on Heterogeneous Mixtures of Particles Model System The sensitivity of the Washburn method was evaluated using single mineral chalcopyritequartz mixtures of varying proportions and known contact angles (See Section for details). The Cassie equation was used to calculate a theoretical contact angle value for each mixture. There is good agreement between the measured and the calculated contact angle for chalcopyrite concentrations above approximately 1% by volume (Figure 6.8). It is evident that the Washburn technique is not sensitive enough to measure the contact angle on particle beds in which there is 1% by volume or less of the hydrophobic component (the measured contact angle for such beds was always zero, as depicted in Figure 6.8). Most natural copper ores have less than 1% chalcopyrite in them, the bulk being hydrophilic 136

159 Advancing contact angle, gangue minerals, so that direct contact angle measurements do not yield meaningful results. In the current study, direct measurements on the feed and tailing size fractions yielded zero contact angle values. However, this result is very encouraging in that for concentrations of the hydrophobic mineral greater than 2%, it should be possible to determine the contact angle on the hydrophobic entity through use of the Cassie equation and armed with assumptions on the contact angle of the gangue component. This methodology is described further below Measured contact angle Cassie Equation Volume fraction of chalcopyrite Figure 6.8 Advancing contact angle as a function of the volume fraction of hydrophobic chalcopyrite in a mixture of clean quartz (zero contact angle) and chalcopyrite µm size fraction for both minerals. It is possible that the detection limit for the hydrophobic component in the chalcopyritequartz mixture may be particle size dependent, and may also be influenced by the particle size in the mixture, and the contact angle of the hydrophobic component. These possibilities require further study. 137

160 Ore System A survey of the mineralogy of the particle size fractions by QEM-Scan shows that the bulk of the chalcopyrite is more than 80% liberated (See Section ; Figure 3.6 for details). In the concentrate, all liberation classes are represented (Figure 6.9), showing that chalcopyrite bearing particles even of very low liberation may be recovered to an extent. This result possibly implies the contact angle of hydrophobic surface is more important in determining flotation behaviour than the degree of liberation, although the latter may influence the former. Particles with contact angle above the critical value for stable bubble particle attachment may be recovered into the concentrate, even though the degree of liberation may be low. The chalcopyrite in the natural ore is associated with other minerals as noted previously. Particle maps show three principal components in the ore feed, i.e., silicates, pyrite and chalcopyrite (Figure 6.10), consistent with calculated results of the composition of the concentrates from assay data in Table 6.2. It is evident from the particle maps that the pyrite that is recovered into the concentrate occurs mainly as fully liberated pyrite or less likely in association with chalcopyrite and/or silicates. This result shows that any contact angle measured on the concentrate is an average of at least three major components, hydrophobic chalcopyrite and pyrite, and hydrophilic quartz. The contact angles measured directly on the concentrate are shown in Table 6.4. The contact angles measured on the concentrate showed small variations, and could be said to be the same, within experimental error, for each size fraction over the range of collector additions. This is because of the presence of hydrophobic pyrite as part of the gangue component. While the grade decreased due to increased mass pull as collector addition increased, the average contact angle remained largely unchanged. This is interpreted as showing that pyrite attains the same contact angle value as chalcopyrite within each size fraction and collector addition. 138

161 Table 6.4 Contact angles measured on the concentrates, the calculated sulphide contact angle values within concentrate, and the back calculated feed contact angles. The back calculated feed contact angles use the R max values. ND not determined. Size fraction, µm Collector addition, g/t Sulphide R max, ±3% Measured contact angle of concentrate, (±2.5 ) Calculated contact angle of sulphide minerals in concentrate, ( ) Back calculated contact angle of sulphide minerals in feed, ( ) ND ND ND ND ND ND ND ND ND ND The contact angle measured on the concentrate size fractions includes contributions from both the hydrophobic minerals and the non-sulphide gangue which is assumed to be hydrophilic. In order to determine the contact angle of chalcopyrite in the concentrate, it was assumed that chalcopyrite and pyrite had the same contact angle within a size fraction and for a given collector addition (i.e. considering sulphur recovery). The non-sulphide gangue component was assumed to have a contact angle of zero (hydrophilic quartz). It is possible that the non-sulphide gangue minerals have a contact angle greater than zero, but the assumption was considered necessary in order to simplify the calculation of the contact angles of the hydrophobic component. The Cassie equation may be used to calculate the mean contact angle when the surface area fractions and contact angles of the components of a mixture are known (Priest et al., 2008). The volume fraction is the same as the surface area fraction for the same size fraction. For a mixture of chalcopyrite, pyrite and nonsulphide gangue, the average contact angle is given by 139

162 (6.1) where θ is the mean contact angle of a given phase. The coefficients x, y, and z, are the respective volume (surface area) fractions of the two minerals. The contact angle of pyrite and chalcopyrite are equal in equation 6.1, and the contact angle of the gangue component is assumed to be zero. Figure 6.9 Liberation of copper sulphides (chalcopyrite) in selected coarse particle size fractions of the concentrate produced with15 g/t DTP. 140

163 Figure 6.10 Particle maps for the ore (a) feed, (b) concentrate, and (c) tailing, for the µm size fraction determined by QEM-Scan. Flotation test at 15 g/t collector addition. NSG is non-sulphide gangue excluding silicates. Cu Sulphides is chalcopyrite. Fe Sulphides is pyrite. Based on knowledge of the composition of each size fraction (Table 6.2), the volume fraction of chalcopyrite, pyrite, and non-sulphide gangue was calculated. The contact angle values of the -20, (2 g/t), and +300 (2 g/t) could not be determined because the Washburn method for contact angle measurements is not sensitive to low concentrations of the hydrophobic component as shown in Figure 6.8. The error in the measurement in these cases was too great. A contact angle of 67 for chalcopyrite particles in the concentrate for the µm size fraction is in good agreement with the value determined indirectly using the ToF-SIMS approach (74±14 ) at 15 g/t DTP. The agreement in the results determined by the two methods is an encouraging development set to lead to new approaches of determining particle wettability, down to the individual particle level. The result confirms the reliability of the Washburn contact angle measurement technique for powder materials. The advancing contact angle values for sulphide minerals in the concentrates for the remainder of the tests are shown in Table 6.4. It must be borne in mind that the contact angle calculated using equation 6.1 and the procedure explained above gives the contact angle of chalcopyrite particles in the concentrate only. As noted before, the calibration curves developed with single mineral chalcopyrite were based on the feed contact angle as the undistributed rate constant is used and its associated collection efficiency. The inferred operational contact angles (Section 141

164 6.2) are based on the feed, and thus, we need to calculate the advancing contact angle of the feed sulphide particles conditioned with collector. To calculate the contact angle of the chalcopyrite in the feed, a back calculation was carried out with the following assumptions: The sulphides in the feed are in two forms, a floatable (R max ) and a non-floatable (100- R max ) component. The latter is locked within gangue and may be assigned a contact angle of zero, and reports to the tailings, even at very high collector additions. The cumulative liberation data for the feed size fractions for chalcopyrite (Figure 3.7) seem to justify this approach. The non-floating component is that part of the sulphides that are lost to the tailing at 30 g/t collector addition and no further collector additions can recover it, as evidenced by the similar recovery at 40 g/t collector addition (Figure 6.1). The difference between the maximum recovery at 30 g/t DTP and the maximum recovery at either 2 or 15 g/t for each size fraction represents that component of the chalcopyrite in the tailing that is recoverable by increasing collector addition. This component requires greater collector to attain a contact angle above the critical value for stable bubble-particle attachment, and was assigned a contact angle equal to the critical value for that size fraction (Table 4.11). This follows from observations from tests with single mineral chalcopyrite that the contact angle of the tailing for each size fraction at its greatest was equal to the critical contact angle, within experimental error (Figure 5.3b). The highest recovery in tests with model chalcopyrite particles was obtained for size fractions between 53 to 150 µm, attaining >98% recovery for a contact angle of 90. In a natural ore, the highest recovery was obtained for the µm size fraction, and decreases as particle size increases. This is considered to be the lowest size fraction for a non-floating component with zero contact angle (i.e., fully locked), above which this component progressively increases. The non-floating component for size fractions above 53 µm was calculated by subtracting the maximum recovery for 30 g/t collector from 100%. This seemed to make sense because a higher collector addition of 40 g/t gave the same recovery as 30 g/t, showing that no additional sulphides could be recovered by increasing the amount of collector. 142

165 The feed contact angle (last column in Table 6.4) was calculated using the Cassie equation. The feed contact angle (back calculated) is the weighted average of the calculated contact angle of the sulphides in the concentrate, the component at the critical contact angle which may report to the tailing at low collector addition, and the component that is locked within gangue, assigned a zero contact angle for all collector additions. The feed contact angles were calculated using the following equations for the 2, 15 and 30 g/t collector additions, respectively, and for each size fraction: (6.2) (6.3) (6.4) where θ c, θ crit are the contact angles of sulphides/chalcopyrite in the concentrate and a component of the tailing with critical contact angle, respectively. R max2, R max15 and R max30 are the fractional maximum recoveries at 2, 15 and 30 g/t, respectively, for each size fraction. The critical contact angles for stable bubble-particle attachment for both chalcopyrite and pyrite have been determined (Figure 5.8). There was a good agreement between the calculated and inferred operational contact angle (Figure 6.11), with an R 2 value of 0.7 across most of the size fractions. Given the complexity of a natural ore with two hydrophobic minerals, the agreement between the inferred operational and measured feed contact angles is acceptable. 143

166 Inferred Contact angle, Calculated Contact angle, Figure 6.11 Comparison of the inferred operational and measured contact angle of sulphide minerals in a natural ore. R 2 = Conclusions The flotation behaviour of particles is closely related to the size and advancing contact angle under constant hydrodynamic conditions. A method to determine an operational contact angle of sulphide mineral particles was developed and uses the undistributed flotation rate constant and collection efficiency to infer the contact angle by benchmarking against a calibration developed with single mineral chalcopyrite. The approach takes into account variations of the contact angle with particle size and infers a contact angle of the feed. This approach has been validated by direct contact angle measurements on combined concentrates on most of the size fractions, and indirectly for the µm size fraction using ToF-SIMS, for a natural porphyry sulphide ore with chalcopyrite and pyrite as the major sulphide minerals. The calibration curves that were generated using single minerals can be used to infer the contact angles of copper minerals in an ore, irrespective of the degree of liberation. It appears the degree of liberation of particles is less important than the contact angle of the chalcopyrite in determining the flotation response of particles for the ore type investigated. 144

167 CHAPTER SEVEN 7. Conclusions Collated data from the flotation of model chalcopyrite particles led to the generation of master curves of the undistributed rate constant and collection efficiency as a function of particle size at different contact angle ranges. These master curves are a useful tool to benchmark the flotation response of an ore. Although the calibration curves have been developed under standardised conditions, the flotation response of an ore, under different hydrodynamic conditions is predictable from theoretical considerations presented in Sections 2.2 and 2.7. Mineral particles achieve different contact angles when they are subjected to the same treatment. Within a size fraction, individual particles have different contact angles with a mean, as measured by the Washburn technique, being the surface area weighted average of the particles. The proportion of the floating and non-floating components in a given size fraction is dependent on the number of particles with contact angle above the critical value required for stable bubble-particle attachment. The critical contact angle varies with particle size fraction. The Washburn technique is a reliable method for the measurement of contact angle on valuable mineral particle beds with a significant hydrophobic component. The method is not sensitive to a low concentration of hydrophobic particles in bulk hydrophilic material. The error in contact angle measurement is small (±2.5 ) and varies with particle size fraction. Generally, the error increases with particle size, a result attributable to the approach to a particle size limit of the method. The advancing contact angle measured using the Washburn technique correlates well with flotation recovery and rate constant. The variation of contact angles with size fraction for similarly treated chalcopyrite particles is not an artefact of the measurement method, but a result of differences in surface 145

168 chemistry, with sulphur species mainly responsible for an increase in contact angle, while oxygen is negatively correlated to hydrophobicity. The distributed and undistributed rate constants converge at high contact angle and diverge as the contact angle decreases due to changes in the non-floating component. The undistributed rate constant is better than the distributed rate constant in characterising flotation response. Statistically, mineral size fractions float in the same manner over a contact angle range of about 5. This contact angle range appears reasonable as it is nearly equal to the error range of the Washburn technique used to measure the contact angle in this study. The size fractions of mineral particles float independently of each other when conditions for first order kinetics are met, which include an excess of the bubble surface, low hydrophobic component in the pulp and no interaction among individual particles. The flotation independence of particles allows the collation and averaging of flotation response for particles of the same size fraction and contact angle range from different tests. Flotation recovery and rate constant increase with an increase in the contact angle of particles. The superficial gas velocity does not affect the maximum recovery, as long as the bubble surface is not limiting, but the rate constant increases with an increase in the superficial gas velocity. As particle size increases from fine to coarse, the flotation rate constant and the recovery increase rapidly, reaching a maximum at the intermediate size fraction, before decreasing with an increase in the particle size. The contact angle of preconditioned feed particles remains relatively constant through flotation with a short conditioning time. Apparently, floatability components in the feed are conserved as particles float with a rate constant that is dependent on the contact angle. Particles that are recovered in the first concentrate have higher contact angles than those in subsequent concentrates. The floating component has a higher contact angle than the nonfloating component (tail) and the latter has a contact angle below the critical for stable bubble-particle attachment. Chalcopyrite particles exhibiting the same contact angle, but with different functional groups on their surfaces floated in a similar manner. Under the conditions used in this study, the flotation response of chalcopyrite particles floated at pulp densities of 2 and 30% w/w (hydrophilic quartz at ph 10) solids was the same. In the flotation of pyrite, the maximum recovery is lower than that for chalcopyrite at similar contact angle, up to 80. The inferred operational contact angle obtained by 146

169 benchmarking against the calibration curves obtained using the chalcopyrite-amyl xanthate system, is lower than the measured feed contact angle across all size fractions. These results are attributed to the higher specific gravity of pyrite and lower stability efficiency. A statistical model fit to pyrite flotation recovery and contact angle (with the same mean and mode contact angle values as for chalcopyrite) resulted in the calculation of critical contact angles, which are higher than for chalcopyrite. The higher critical contact angle may be the reason for the lower recovery below 80 and the lower inferred operational contact angle. The calibration curves appear therefore, to be mineral specific. The flotation of a natural copper ore showed that recovery and rate constant increased with an increase in collector addition. The increase in collector addition is equivalent to an increase in the contact angle. Apparently, fine particles adsorb collector in preference to the coarse particles, the latter benefitting more from higher collector additions. The benchmarking of the flotation response of a porphyry copper ore against the calibration curves produced reasonable agreement between the calculated feed and inferred operational contact angles, after a correction for unliberated sulphide particles for size fractions above 53 µm. A technique for the measurement of contact angle of particles based on their flotation response has been developed Future Work Opportunities to strengthen the methodology developed in this thesis abound. There is need to test and quantify the influence of particle size distribution on the flotation response of minerals, and possibly establish boundaries of applicability with respect to the proportion of fine and coarse particles in the feed. Concentrates collected at different times during flotation show different contact angles. Further study could explore the possibility of estimating contact angle distribution of particles in the feed from the recovery with time profile of each size fraction. The statistical distribution of contact angles around the mean may need further study to establish whether the behaviour varies with particle size fraction, mineral surface species or mineral type. The critical contact angle for stable bubble-particle attachment depends on size and mineral specific gravity under constant hydrodynamic conditions. A study to quantify the change in the critical contact angle with a change in specific gravity would greatly expand 147

170 the applicability of the calibration curves generated in this study, as it would lead to the modelling of the maximum recovery of any mineral destined for separation by flotation. 148

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183 Appendices 161

184 APPENDIX 1 A1 Material Characterisation A1.1. Chemical assay of chalcopyrite Chalcopyrite as supplied was significantly contaminated with silica (8%) on an un-sized basis. Collectorless flotation was successful in rejecting much of the silica contamination (Table A1.1). Table A1.1 Chalcopyrite assay, wt%, before (B) and after (A) a pre-flotation cleaning flotation stage B A B A B A B A B A B A B A B A B A Fe Cu S SiO MgO CaO A1.2. Scanning Electron Microscope Images of Chalcopyrite Size Fractions 162

185 Figure A1.1 SEM microgram of -20 µm chalcopyrite size fraction Figure A1.2 SEM microgram of µm chalcopyrite size fraction 163

186 Figure A1.3 SEM microgram of µm chalcopyrite size fraction Figure A1.4 SEM microgram of µm chalcopyrite size fraction 164

187 Figure A1.5 SEM microgram of µm chalcopyrite size fraction Figure A1.6 SEM microgram of µm chalcopyrite size fraction 165

188 Figure A1.7 SEM microgram of µm chalcopyrite size fraction Figure A1.8 SEM microgram of µm chalcopyrite size fraction 166

189 Figure A1.9 SEM microgram of +300 µm chalcopyrite size fraction 167

190 APPENDIX 2 A2 Data Used in Error Analysis And Surface Analysis A2.1. Contact Angle Measurements on Sized And Homogenised Chalcopyrite Samples For each size fraction, 150 g of chalcopyrite was conditioned with collector using the procedure outlined in Chapter 3. Five pairs of 2 g samples each were sampled from each size fraction for contact angle measurements. Three flotation tests (feed size distribution as outlined in Chapter 3, and each test repeated twice) were subsequently carried out to obtain the flotation response corresponding to the measured mean contact angle of each size fraction. Table A2.1 Repeated Contact angle measured on homogenised chalcopyrite sample (conditioned with 200 g/t potassium amyl xanthate), high contact angle. Size Fraction, µm Measurement Number Mean Contact Angle, ( ) Table A2.2 Maximum recovery, (%) and undistributed rate constant (min -1 ) for homogenised sample with contact angle in Table A2.1. Size Fraction, µm Paramet er Test R m ax Mean k* R m ax k* R m ax k* R m ax k* 168 R m ax k* R m ax k* R m ax k* R m ax k* R m ax k*

191 Table A2.3 Repeated Contact angle measured on homogenised chalcopyrite sample (conditioned with 50 g/t potassium amyl xanthate), Low contact angle. Size Fraction, µm Measurement Number Mean Contact Angle, ( ) Table A2.4 Maximum recovery, (%) and undistributed rate constant (min -1 ) for homogenised sample with contact angle in Table A2.3. Size Fraction, µm Paramet er Test Mean R ma x k* R ma x k* R ma x k* R ma x k* R ma x k* R ma x k* R ma x k* R ma x k* R ma x k* A2.2. Surface Analysis Contact angle measurements on collectorless chalcopyrite size fractions were carried out in conjunction with surface analysis by XPS in order to establish the correlation between contact angle and surface species. The contact angle for each size fraction in Table A2.5 is correlated to surface species in Figures Table A2.5 Advancing contact angle of collectorless chalcopyrite size fractions Size Fraction, µm Contact Angle, ( )

192 Figure A2.1 XPS spectra for -20 µm size fraction, surface atomic concentrations. Figure A2.2 XPS spectra for -20 µm size fraction, Cu species 170

193 Figure A2.3 XPS spectra for -20 µm size fraction, Fe species Figure A2.4 XPS spectra for -20 µm size fraction, O species 171

194 Figure A2.5 XPS spectra for -20 µm size fraction, C species Figure A2.6 XPS spectra for -20 µm size fraction, S species 172

195 Figure A2.7 XPS spectra for µm size fraction, surface atomic concentration. Figure A2.8 XPS spectra for µm size fraction, Cu species. 173

196 Figure A2.9 XPS spectra for µm size fraction, Fe species. Figure A2.10 XPS spectra for µm size fraction, O species. 174

197 Figure A2.11 XPS spectra for µm size fraction, C species. Figure A2.12 XPS spectra for µm size fraction, S species. 175

198 Figure A2.13 XPS spectra for µm size fraction, surface atomic concentration. Figure A2.14 XPS spectra for µm size fraction, Cu species. 176

199 Figure A2.15 XPS spectra for µm size fraction, Fe species. Figure A2.16 XPS spectra for µm size fraction, O species. 177

200 Figure A2.17 XPS spectra for µm size fraction, C species. Figure A2.18 XPS spectra for µm size fraction, S species. 178

201 Figure A2.19 XPS spectra for µm size fraction, surface atomic concentration. Figure A2.20 XPS spectra for µm size fraction, Cu species. 179

202 Figure A2.21 XPS spectra for µm size fraction, Fe species. Figure A2.22 XPS spectra for µm size fraction, O species. 180

203 Figure A2.23 XPS spectra for µm size fraction, C species. Figure A2.24 XPS spectra for µm size fraction, S species. 181

204 Figure A2.25 XPS spectra for µm size fraction, surface atomic concentration. Figure A2.26 XPS spectra for µm size fraction, Cu species. 182

205 Figure A2.27 XPS spectra for µm size fraction, Fe species. Figure A2.28 XPS spectra for µm size fraction, O species. 183

206 Figure A2.29 XPS spectra for µm size fraction, C species. Figure A2.30 XPS spectra for µm size fraction, S species. 184

207 Figure A2.31 XPS spectra for µm size fraction, surface atomic concentration. Figure A2.32 XPS spectra for µm size fraction, Cu species. 185

208 Figure A2.33 XPS spectra for µm size fraction, Fe species. Figure A2.34 XPS spectra for µm size fraction, O species. 186

209 Figure A2.35 XPS spectra for µm size fraction, C species. Figure A2.36 XPS spectra for µm size fraction, S species. 187

210 Figure A2.37 XPS spectra for µm size fraction, surface atomic concentration. Figure A2.38 XPS spectra for µm size fraction, Cu species. 188

211 Figure A2.39 XPS spectra for µm size fraction, Fe species. Figure A2.40 XPS spectra for µm size fraction, O species. 189