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Flocculation, hydrophobic agglomeration and filtration of ultrafine coal Yu, Zhimin 1999

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FLOCCULATION, HYDROPHOBIC AGGLOMERATION AND FILTRATION OF ULTRAFINE COAL by Zhimin Yu B . A . S c , Huainan Institute of Mining and Technology, 1982 M . A . S c , China University of Mining and Technology, 1987 A THESIS SUBMITTED IN PARTIAL FULFILLMENT OF THE REQUIREMENTS FOR THE D E G R E E OF DOCTOR OF PHILOSOPHY In THE FACULTY OF GRADUATE STUDIES Department of Mining and Mineral Process Engineering We accept this thesis as conforming to the required standard THE UNIVERSITY OF BRITISH COLUMBIA Zhimin Yu, 1998 In presenting this thesis in partial fulfilment of the requirements for an advanced degree at the University of British Columbia, I agree that the Library shall make it freely available for reference and study. I further agree that permission for extensive copying of this thesis for scholarly purposes may be granted by the head of my department or by his or her representatives. It is understood that copying or publication of this thesis for financial gain shall not be allowed without my written permission. Department p i / M ^ ^ S &^Mt^J The University of British Columbia Vancouver, Canada Date L Z ^ . / DE-6 (2/88) 11 A B S T R A C T In coal preparation plant circuits, fine coal particles are aggregated either by oil agglomeration or by flocculation. In a new hydrophobic agglomeration process, recently developed hydrophobic latices are utilized. While the selectivity of such aggregation processes determines the beneficiation results, the degree of aggregation has a strong effect on fine coal filtration. The aim of this research was to study the fundamentals and analyze the common grounds for these processes, including the potential effect of the coal surface properties. The selective flocculation tests, in which three types of coal, which differed widely in surface wettability, and three additives (hydrophobic latices, a semi-hydrophobic flocculant and a typical hydrophilic polyelectrolyte) were utilized, showed that coal wettability plays a very important role in selective flocculation. The abstraction of a hydrophobic latex on coal and silica revealed that the latex had a much higher affinity towards hydrophobic coal than to hydrophilic mineral matter. As a result, the U B C - 1 hydrophobic latex flocculated only hydrophobic coal particles while the polyeletrolyte (PAM) flocculated all the tested coal samples and minerals, showing no selectivity in the fine coal beneficiation. The oil agglomeration was tested using kerosene emulsified with various surfactants (e.g. cationic, anionic and non-ionic). Surfactants enhance not only oil emulsification, hence reducing oil consumption (down to 0.25-0.5%), but also entirely change the electrokinetic properties of the droplets and affect the interaction energy between oil droplets and coal particles. Consequently, the results found in the course of the experimental work strongly indicate that even oxidized coals can be agglomerated if cationic surfactants are used to emulsify the oil. Oil agglomeration of the Ford-4 ultrafine coal showed that even at extremely low oil consumption (0.25 to 0.5%), a clean coal product with an ash content around 5% at over 99.9% coal recovery could be obtained in a one-stage separation by screening the agglomerated product. If a conventional oil agglomeration process is used instead, oil consumption as high as 30% is needed to obtain comparable results. In the tests on filtration and dewatering of ultrafine and fine coals, the effect of chemical additives and coal surface properties was investigated. The tests revealed very significant differences in the filtration of ultrafine (-45 um) and fine (-500 um) coals. The moisture contents in the filter cakes in the tests with ultrafine coal were around 40% (irrespective of the coal surface properties), while for the fine coal the moisture content fluctuated around 18% (Ford-4) and 30% (Ford-13). The results revealed that the hydrophobic latex and the emulsified oils could not only successfully beneficiate the ultrafine coal but also significantly increase filtration rate and/or reduce moisture content of the filter cake. Among the chemicals tested, the emulsified oils were found to be the most promising not only for the beneficiation but also for filtration and dewatering processes. Surfactants were found to only slightly affect the filtration of fine coal. However, they can influence filtration very profoundly if utilized to emulsify the oil which is used to agglomerate coal prior to its filtration. I l l FLOCCULATION, HYDROPHOBIC AGGLOMERATION AND FILTRATION OF ULTRAFINE COAL TABLE OF CONTENTS ABSTRACT ii T A B L E OF CONTENTS i i i LIST OF FIGURES ix LIST OF TABLES xvii NOMENCLATURE xviii ACKNOWLEDGMENTS xx CHAPTER 1. INTRODUCTION 1 CHAPTER 2. RESEARCH OBJECTIVES 6 CHAPTER 3. LITERATURE REVIEW 8 3.1 Surface Properties of Coal ....8 3.1.1 Coal rank and chemical composition 8 3.1.2 Porosity and surface area 10 3.1.3 Functional groups 13 3.1.4 Surface electrical charge 17 3.2 Interaction of Chemical Reagents with Coal Surface 23 3.2.1 Stability of coal suspension and effect of electrolytes or flocculants 23 3.2.2 Selective flocculation and hydrophobic agglomeration 27 3.2.3 Interaction of surfactants with solid surface 42 3.2.3.1 Interaction of surfactants with solid surface 42 iv 3.2.3.2 Adsorption mechanism 44 3.3 Filtration and Dewatering of Fine and Ultrafine Coal 47 3.3.1 Theory of filtration and dewatering 48 3.3.1.1 Theory for filtration process under constant pressure 50 3.3.1.2 Theory for dewatering 52 3.3.1.2.1 Model for dewatering 53 3.3.1.2.2 Model for residual saturation of the cake...54 3.3.1.3 Enhancement of filtration and dewatering 55 3.3.2 Filtration enhanced by flocculants 57 3.3.3 Filtration enhanced by surfactants 61 3.3.4 Filtration enhanced by hydrophobization 68 C H A P T E R 4. M A T E R I A L S 73 4.1 Coal Samples 73 4.1.1 Coal samples preparation for filtration tests 73 4.1.2 Coal sample preparation for hydrophobic agglomeration of ultrafine coal 75 4.1.3 Sample preparation for abstraction of latex and oil droplets 76 4.2 Silica Sample 76 4.3 Clay Sample 76 4.4 Chemical Reagents 77 4.4.1 Polyelectrolytes 77 4.4.2 Hydrophobic latices 77 4.4.3 Semi4iydrophobic flocculant 78 4.4.4 Surfactants 78 4.4.4.1 Cationic surfactants 78 4.4.4.2 Anionic surfactants 79 4.4.4.3 Non-ionic surfactants 79 4.4.5 Oily hydrocarbons 80 4.4.6 Dispersing agents 80 4.4.7 pH regulators 81 CHAPTER 5. EXPERIMENTAL TECHNIQUES 82 5.1 Adsorption and Attachment Measurements 82 5.1.1 Abstraction of a hydrophobic latex by coal particles 82 5.1.2 Determination of hydrophobic latex concentration 83 5.1.3 Adsorption of surfactants 85 5.1.4 Abstraction of oil droplets by coal particles 89 5.2 Emulsification of Oily Hydrocarbons 91 5.3 Assessment of Coal Surface Wettability 93 5.4 Measurement of Particle Size Distributions 93 5.5 Electrokinetic Tests 94 5.6 Flocculation and Hydrophobic Agglomeration 95 5.7 Oil Agglomeration 96 5.8 Filtration and Dewatering 97 5.9 Surface area Determination 98 5.10 Defination of Measurements in Beneficiation Tests 99 CHAPTER 6. RESULTS AND DISCUSSION 100 6.0 Introduction 100 6.1 Total and Selctive Flocculation of Ultrafine Coal 101 6.1.1 Abstraction of a hydrophobic latex by coal particles 101 vi 6.12 Effect of flocculant type and coal wettability on flocculation and hydrophobic agglomeration 104 6.1.3 Effect of p H on flocculation and hydrophobic agglomeration 108 6.1.4 Effect of hydrodynamic conditions 110 6.15 Summary and discussion 110 6.2 Beneficiation of Ultrafine Coal using Hydrophobic Latices 115 6.2.1 Predetermination of slurry solids content for selective flocculation 115 6.2.2 Effect of dispersant on selective flocculation 116 6.2.3 Separation of coal from mixture of coal-silica by hydrophobic agglomeration 118 6.2.4 Separation of coal and kaolin by hydrophobic agglomeration 123 6.2.5 Beneficiation of ultrafine run-of-mine coal by hydrophobic agglomeration 127 6.2.6 Summary and discussion 129 6.3 Oi l Agglomeration of Ultrafine Coal with Emulsified Oils 131 6.3.1 Emulsification of liquid hydrocarbons 131 6.3.1.1 O i l droplet size distribution 131 6.3.1.2 Electrokinetic potentials of oil-droplets in aqueous solutions 133 6.3.2 Abstraction of oil-droplets by coal particles 138 6.3.3 Beneficiation of ultrafine coal with emulsified oil 142 6.3.3.1 Oil agglomeration of ultrafine coal with cationic emulsions 142 Vll 6.3.3.2 Oi l agglomeration of ultrafine coal with anionic emulsion 146 6.3.3.3 Beneficiation of ultrafine coal with non-ionic emulsion 150 6.3.3.4 Comparison of beneficiation results using kerosene emulsified with different surfactants 151 6.3.4 Summary and discussion 153 Dewatering of Ultrafine Coal 162 6.4.1 Parameters influencing filtration of ultrafine coal 162 6.4.2 Filtration with the use of flocculants including polyelectrolytes, semi-hydrophobic flocculants and hydrophobic latices 163 6.4.2.1 Effect of polyelectrolytes and hydrophobic latices on coal surface wettability... 163 6.4.2.2 Filtration of hydrophobic metallurgical coal and semi-hydrophobic coal with three different types of flocculants 164 6.4.2.3 Filtration of hydrophilic oxidized coal with three types of additives 171 6.4.3 Filtration with the use of surfactants and emulsified oils 173 6.4.3.1 Adsorption of surfactants on coal surface 174 6.4.3.2 Filtration of ultrafine hydrophobic coal with surfactants and oil emulsified oils 180 6.4.3.3 Filtration of hydrophilic ultrafine coal with surfactants and emulsified oils 186 6.4.4 Effect of particle size on filtration with various additives 190 6.4.4.1 Filtration of -500 um hydrophobic coal with flocculants, surfactants and emulsified oils 190 6.4.4.2 Filtration of -500 um hydrophilic coal with surfactants and emulsified oils 192 J 6.4.5 Filtration of ultrafine coal under high pressure 193 6.4.6 Summary and discussion. 194 CHAPTER 7. CONCLUSIONS AND RECOMMENDATIONS 203 REFERENCES 210 APPENDIX 225 ix L I S T O F F I G U R E S Figure 3.1.1-1 Carbon distribution in coals of varying rank 9 Figure 3.1.3-1. Functional groups in coal 15 Figure 3.1.3-2. Schematic representation of possible coal surface polar groups 15 Figure 3.1.4-1 Schematic representation of an electric double layer 18 Figure 3.1.4-2 Schematic representation of a fiffuse electric double layer 19 Figure 3.1.4-3. Generalized zeta potential versus p H diagram for coals of various ranks20 Figure 3.2.1-1. Potential energy curves 24 Figure 3.2.3.1-1 Model for adsorption of surfactant molecules onto hydrophobic and hydrophilic solid surfaces 43 Figure 3.2.3.2-1 Adsorption of isotherms for sodium alkylsulfonates of different hydrocarbon chain lengths on alumina at 2x10 3 mol/1 ionic strength 46 Figure 3.2.3.2-2 The sessile drop contact angles for a HVA-bituminous coal and an anthracite coal as function of Triton N - l 01 concentration 47 Figure 3.3.3-2 The models for adsorption of anionic and cationic surfactants on coal ...66 Figure 3.3.4-1 Effect of oil concentration on agglomerate size and moisture content 71 Figure 4.1.1-1 Zeta potential vs. p H curves for the tested coal samples 75 Figure 5.1.2-1 Calibration curve of UBC-1 latex concentration versus solution transimittance (The solution transmittance was measured at a light wavelength of 800 nm) '. 84 Figure 5.1.3-1 Sketch of experimental set-up for measurement of surfactant concentration with the use of ion-surfactant selective electrodes 87 Figure 5.1.3-2 Calibration curve showing electrode potential versus concentration of D D A . H C 1 88 Figure 5.1.3-3 Calibration curve showing electrode potential vs. concentration of DDS.88 Figure 5.1.4-1 Calibration curve of kero-DDA oil emulsion concentration vs. transmittance 90 Figure 5.1.4-2 Calibration curve of kero-DDS oil emulsion concentration vs. transmittance 91 Figure 6.1.1-1 Abstraction kinetics of U B C - 1 latex by Ford-4 coal and silica particles at p H 6.4 102 Figure 6.1.1-2 Abstraction isotherms of U B C - 1 on Ford-4 coal and silica at p H 6.4....102 Figure 6.1.1-3 The effect of p H on the abstraction of UBC-1 latex by Ford-4 coal and silica particles 103 Figure 6.1.2-1 Effect of U B C - 1 latex dosage and coal wettability on hydrophobic flocculation of ultrafine coal at p H 6.8 105 Figure 6.1.2-2 Effect of P E O dosage and coal wettability on flocculation of ultrafine coal a tpH6.8 105 Figure 6.1.2-3 Effect of P A M dosage and coal wettability on flocculation of ulrafine coal at p H 6.8 106 Figure 6.1.2-4 Comparison of U B C - 1 with F R - 7 A latex at p H 6.8 106 Figure 6.1.3-1 Effect of p H on flocculation and hydrophobic agglomeration (UBC-1 dosage: 200 g/t, P A M dosage: 200 g/t) 109 Figure 6.1.3-2 Effect of hydrodynamic conditions on flocculation and hydrophobic agglomeration (UBC-1 dosage: 200 g/t, P A M dosage: 200 g/t, Ford-4 ultrafine coal) 109 Figure 6.1.5-1 Zeta potential of coal particles and U B C - 1 latex particles, and hydrophobic agglomeration of Ford-4 and Ford-13 coals using U B C - 1 latex 113 Figure 6.2.1-1 The effect of solids content on hydrophobic agglomeration 115 Figure 6.2.2-1 Dispersion tests with sodium hexametaphosphate using ultrafine Ford-4 coal 116 Figure 6.2.2-2 Effect of dispersant (SHMP) dosage on hydrophobic agglomeration of a mixture of 90% ultrafine coal and 10% ultrafine kaolin with U B C - 1 latex (UBC-1 dosage: 400 g/t, natural p H 6.4, shear rate: 320 rpm, Ford-4 sample) 117 Figure 6.2.3-1 Selective flocculation tests of a 50:50% coal/silica mixture as a function of flocculant dosage 120 xi Figure 6.2.3-2 Selective flocculation tests of a 50:50% coal/silica mixture as a function o f p H . 121 Figure 6.2.3-3 Selective flocculation tests of a 50:50%) coal/silica mixture as a function of stirring rate (rpm) 122 Figure 6.2.4-1. Effect of UBC-1 dosage on selective flocculation of coal-kaolin mixture (Natural p H 6.2-6.4, stirring rate 320 rpm, S H M P dosage 200 mg/1) 124 Figure 6.2.4-2. Effect of p H on selective flocculation of coal-kaolin mixtures (UBC-1 dosage 300 g/t, stirring rate 320 rpm, S H M P dosage 200 mg/1) 125 Figure 6.2.4-3. The effect of stirring rate on selective flocculation of coal-kaolin mixtures (UBC-1 dosage 300 g/t, natural pH6.2-6.4, S H M P dosage 200 mg/1) 126 Figure 6.2.5-1 Effect of U B C - 1 dosage on cleaning of ultrafine Ford-4 coal by selective flocculation (Natural pH6.7, stirring rate 320 rpm, S H M P 300 mg/1) 128 Figure 6.2.5-2 Effect of p H on cleaning of ultrafine Ford-4 coal by selective flocculation (UBC-1 dosage 400 g/t, stirring rate 320 rpm, S H M P 300 mg/1) 128 Figure 6.2.5-3 Effect of stirring rate on cleaning of finely ground Ford-4 coal by selective flocculation (Natural p H 6.7, UBC-1 dosage 400 g/t, S H M P 300 mg/1) .' 129 Figure 6.3.1.2-1 Zeta potential of the kerosene droplets emulsified in distilled water and in 5.4 x 10-3 , 5.4 x 10"4 and 5.4 x 10"5 M dodecyl amine solutions 134 Figure 6.3.1.2-2 Zeta potential of the kerosene droplets emulsified in distilled water and in 3.9 x 10'4 M and 3.9 x 10-5 M C 1 6 . l g long chain amine solutions 135 Figure 6.3.1.2-3 Zeta potential of the kerosene droplets and kerosene containing 1% and 10%) D D A emulsified in distilled water 136 Figure 6.3.1.2-4 Zeta potential of the kerosene droplets and kerosene containing 1% and 0.1%) C 1 6 . 1 8 amine emulsified in distiied water 136 Figure 6.3.1.2-5 Zeta potential of the oil droplets emulsified in distilled water and in 2.7 x 10~4 M sodium cetyl sulfate and 3.4 x 10~3 M dodecyl sulfate solutions 137 \ Xll Figure 6.3.1.2-6 Zeta potential of oil droplets emulsified in distilled water and in 3.7 x 10-4 M CO-520 and 1.7 x 1 0 4 M CO-610 solutions 137 Figure 6.3.2-1 Abstraction kinetics of Kero-DDA oil droplets by Ford-4 coal (Coal size 0.6x0.045 mm, oil droplets concentration 200 mg/1, 350 rpm, natural p H (6.4)) 139 Figure 6.3.2-2 Abstraction kinetics of Kero-DDS oil droplets by Ford-4 coal (Coal Size 0.6x0.045 mm, oil droplets concentration 200 mg/1, 350 rpm, natural p H (6.3)) 139 Figure 6.3.2-3 Abstraction of K e r o - D D A oil droplets by Ford-4 coal (Coal size 0.6x0.045 mm, 20 grams, abstraction time 15 minutes, 350 rpm, natural p H (6.4), content of D D A in kerosene 1 wt%) 140 Figure 6.3.2-4 Abstraction of Kero-DDS oil droplets by Ford-4 coal (Coal size 0.6x0.045 mm, 20 grams, abstraction time 15 minutes, 350 rpm, natural p H 6.4, content of D D S in kerosene 1 wt%)) 140 Figure 6.3.3.1-1 Agglomeration of ultrafine Ford-4 raw coal with the use of kero-DDA (Test conditions: p H 7.8, 20,000 rpm, S H M P 100 ppm) 143 Figure 6.3.3.1-2 The effect of p H on agglomeration of ultrafine Ford-4 raw coal using cationic kero-DDA emulsion (Test conditions: oil dosage 2%, 20,000 rpm, S H M P 100 ppm) 144 Figure 6.3.3.1-3 The effect of stirring rate on agglomeration of ultrafine Ford-4 raw coal with the use of cationic kero-DDA emulsion (Test conditions: oil dosage 2%, pH7.8, S H M P 100 ppm) 145 Figure 6.3.3.2-1 Agglomeration of ultrafine Ford-4 raw coal with the use of a kero-DDS anionic emulsion (Test conditions: pH7.7 , 20,000 rpm, S H M P 100 ppm) 147 Figure 6.3.3.2-2 The effect of p H on agglomeration of ultrafine Ford-4 raw coal using anioinc kero-DDS oil emulsion (Test conditions: oil dosage 2%, 20,000 rpm, S H M P 100 ppm) 148 Xlll Figure 6.3.3.2-3 The effect of stirring rate on agglomeration of ultrafine Ford-4 raw coal with the addition of anionic kero-DDS oil emulsion (Test conditions: oil dosage 2%, p H 7.7, S H M P 100 ppm) 149 Figure 6.3.3.3-1 Agglomeration of ultrafine Ford-4 coal using kero-CO-610 emulsionl50 Figure 6.3.3.3-2 Agglomeration of ultrafine Ford-4 coal using kero-CO520 emulsion. 150 Figure 6.3.3.3-4 Comparison of oil agglomeration using emulsified oil with conventional oil agglomeration and selective hydrophobic agglomeration using the same ultrafine coal (Test conditions: Ford-4 ultrafine raw coal, p H 7-8, oil agglomeration stirring rate 20,000 rpm, flocculation stirring rate 300 rpm, S H M P in agglomeration 100 ppm, S H M P in selective flocculation 300 ppm) 152 Figure 6.3.4-1 Schematic comparison of the aggloeration using emulsified oils with a conventional oil agglomeration 157 Figure 6.3.4-2 Zeta potentials of the emulsified oil droplets and solid particles and corresponding oil agglomeration results 159 Figure 6.4.2.1-1 The effect of flocculant type on L C - 7 coal surface wettability 164 Figure 6.4.2.2-1 The effect of flocculant type and dosage on filtration rate of Ford-4 ultrafine coal 165 Figure 6.4.2.2-2 The effect of the flocculant type and dosage on cake moisture content of Ford-4 ultrafine coal 165 Figure 6.4.2.2-3 The effect of flocculant type and dosage on filtration rate of L C - 7 ultrafine coal 167 Figure 6.4.2.2-4 The effect of flocculant type and dosage on cake moisture content of L C - 7 ultrafine coal 167 Figure 6.4.2.2-5 The effect of stirring rate and flocculant type on filtration rate of hydrophobic Ford-4 ultrafine coal 168 Figure 6.4.2.2-6 The effect of shear rate and flocculant type on cake moisture content of hydrophobic Ford-4 ultrafine coal 168 Figure 6.4.2.2-7 The effect of shear rate and flocculant type on filtration rate of L C - 7 ultrafine coal 169 xiv Figure 6.4.2.2-8 The effect of shear rate and flocculant type on cake moisture content of L C - 7 ultrafine coal 169 Figure 6.4.2.2-9 The effect of p H on filtration rate of Ford-4 ultrafine coal 170 Figure 6.4.2.2-10 The effect of p H on cake moisture of Ford-4 ultrafine coal 170 Figure 6.4.2.3-1 The effect of flocculant type and dosage on filtration rate of hydrophilic Ford-13 ultrafine coal 171 Figure 6.4.2.3-2 The effect of flocculant type and dosage on cake moisture content of Ford-13 ultrafine coal 172 Figure 6.4.2.3-3 The effect of stirring rate and flocculant type on filtration rate of Ford-13 ultrafine coal 172 Figure 6.4.2.3-4 The effect of stirring rate and flocculant type on cake moisture content of Ford-13 ultrafine coal 173 Figure 6.4.3.1-3 Adsorption of dodecyl amine on ultrafine Ford-4 coal at p H 5.92-5.93 175 Figure 6.4.3.1-4 Adsorption of sodium dodecyl sulfate on ultrafine Ford-4 coal at p H 5.67-5.68 175 Figure 6.4.3.2-1 The effect of oil surfactants and emulsified oils on filtration rate of ultrafine Ford-4 coal 181 Figure 6.4.3.2-2 The effect of surfactants and emulsified oils on cake moisture content of ultrafine Ford-4 coal 181 Figure 6.4.3.2-3 Filtration rate of Ford-4 ultrafine coal using various anionic oil emulsions 182 Figure 6.4.3.2-4 Cake moisture content of Ford-4 ultrafine using anionic emulsions.. 182 Figure 6.4.3.2-5 Filtration rate of Ford-4 ultrafine coal using cationic oil emulsions.... 184 Figure 6.4.3.2-6 Cake moisture content of Ford-4 ultrafine using cationic oil emulsions 184 Figure 6.4.3.2-7 The effect of p H on filtration rate of Ford-4 ultrafine coal with the use of kerosene emulsified with cationic and anionic surfactants (DDA-1: kero emulsified in 5.4xl0"4 M D D A solution, D D A - 5 : kero containing 1% D D A emulsified in distilled water, C16-18-4: kero containing 1% C16-18 XV amine emulsified in distilled water, SCTS-1: kero emulsified in 2.7x10"4 M S C T S solution, DDS: kero emulsified in 3.4X10"4 M SDDS solution) 185 Figure 6.4.3.2-8 The effect of p H on cake moisture content of Ford-4 ultrafine coal with the use of kerosene emulsified with cationic and anionic surfactants 185 Figure 6.4.3.3-1 The effect of surfactants and oil emulsions on filtration rate of hydrophilic Ford-13 ultrafine coal 186 Figure 6.4.3.3-2 The effect of surfactants and oil emulsions on cake moisture content of hydrophilic Ford-13 ultrafine coal 187 Figure 6.4.3.3-3 Filtration rate of hydrophilic Ford-13 ultrafine coal with the use of kerosene emulsified with various surfactants 189 Figure 6.4.3.3-4 Cake moisture content of hydrophilic Ford-13 ultrafine coal with the use of kerosene emulsified with various surfactants 189 Figure 6.4.4.1-1 The effect of flocculants, surfactants and oil emulsions one filtration rate of -0.5 mm hydrophobic Ford-4 fine coal 191 Figure 6.4.4.1-2 The effect of flocculants, surfactants and oil emulsions on cake moisture content of-0.5 mm hydrophobic Ford-4 fine coal 191 Figure 6.4.4.2-1 The effect of flocculants, surfactants and oil emulsions on filtration rate of -0.5 mm hydrophilic Ford-13 fine coal 192 Figure 6.4.4.2-2 The effect of flocculants, surfactants and oil emulsions on cake moisture content of -0.5 mm hydrophilic Ford-13 fine coal 193 Figure 6.4.5-1 The effect of filtration pressure on cake moisture content of ultrafine Ford-4 coal with addition of hydrophobic latex F R - 7 A 194 Figure 6.4.6-1 The structure of filter bed formed by ultrafine and fine coal with addition of flocculants or other additives 196 Figure6.4.6-2 The zeta-potentials of Ford-4 coal particles and oil droplets emulsified with cationic and anionic surfactants 199 Figure Apdx-1 The filter bed configurations during filtration and dewatering stages..227 Figure Apdx-2 Zeta potentials of UBC-1 latex particles (Appendix) 230 Figure Apdx-3 Flocculation and filtration test procedures (Appendix) 231 xvi Figure Apdx-4 Sketch of filtration and dewatering system (Appendix) 232 Figure Apdx-5 Size distributions of kerosene droplets emulsified in a 5.4x10^ M dodecyl amine solution 234 Figure Apdx-6 Size distributions of kerosene droplets obtained by emulsifying kerosene containing lwt% of dodecyl amine emulsified in aqueous solution 234 Figure Apdx-7 Size distributions of kerosene droplets emulsified in a 5.4xl0 - 5 M dodecyl amine solution 235 Figure Apdx-8 Size distributions of kerosene droplets obtained by emulsifying kerosene containing 0.1 wt% of dodecyl amine in aqueous solution 235 Figure Apdx-9 Size distributions of kerosene droplets emulsified in a 3.9x10"4 M C 1 6 . 1 8 amine solution 236 Figure Apdx-10 Size distributions of kerosene droplets obtained by emulsifying kerosene containing lwt% of C 1 6 . ] 8 amine in aqueous solution 236 Figure Apdx-11 Size distributions of kerosene droplets emulsified in a 3.9xl0"5 M C 1 6 . l g amine solution 237 Figure Apdx-12 Size distributions of kerosene droplets obtained by emulsifying kerosene containing 0.1 wt% of C 1 6 . 1 8 amine in aqueous solution 237 Figure Apdx-13 Size distributions of kerosene droplets emulsified in a 3.4xl0"4 M dodecyl sulfate solution 238 Figure Apdx-14 Size distributions of kerosene droplets emulsified in a 2.7xl0 - 4 M cetyl sulfate solution 238 Figure Apdx-15 Size distributions of kerosene droplets emulsified in a 1.7xl0"4 M nonylphenoxy polyethanol CO-610 solution 239 Figure Apdx-16 Calculation of surface area of Ford-4 ultrafine coal based on the data shown in Table Apdx-8 239 Figure Apedx-17 Typical flowsheet of refinery of E S S O oils (Appendix) 244 xvn L I S T O F T A B L E S Table 4.1-1 Assessment of coal surface wettability and other related properties 73 Table 4.1.1-1. Proximate Analyses of Ford-4, Ford-13 and Line Creek-7 Coal Samples.74 Table 6.4.3.1-1 Aggregation of ultrafine coal using D D A and D D S (natural p H 6.4)...178 Table Apdx-1. Ford-4, Ford-13 and L C - 7 Ultrafine Coal Sample Size Distributions (Appendix) 226 Table Apdx-2. Ford-4 Fine Coal Particle Size Distribution (Appendix) 227 Table Apdx-3. Ford-13 Fine Coal Particle Size Distribution (Appendix) 228 Table Apdx-4. Size Distribution of Ford-4 Ultrafine Coal by Wet Grinding (Appendix) 228 Table Apdx-5. The size distribution of coal sample used in abstraction tests (Appendix) 229 Table Apdx-6. The size distribution of silica sample used in abstraction tests (Appendix) 229 Table Apdx-7. The Size of Oil-droplets in Various Emulsions (Appendix) 233 Table Apdx-8. Calculation of surface area of the Ford-4 ultrafine coal 240 Table Apdx-9. Electrode potential measurements dodecyl amine chloride solutions at different concentrations using the selective electrode (Appendix) 241 Table Apdx-10. The data for calibration curve of U B C - 1 latex concentration against solution transmittance at light wavelength of 800 nm (Appendix) 241 Table Apdx-11 The data for calibration curve of kero-DDA emulsion oil concentration against solution transmittance at light wavelength of 800 nm (Appendix) 242 Table Apdx-12 Size distribution measurements for kerosene droplets emulsified in a 2.7x10"4 M cetyl sulfate solution (Appendix) 242 Table Apdx-13. Measurements of filtration rate and cake moisture contents of Ford-4 ultrafine coal with the use of different additives (Appendix) 243 Table Apdx-14 Mass spectrometer analysis - oil composition and physical-chemical inspections of the oil components 245 N O M E N C L A T U R E A - cross-sectional area of the filter bed Af - feed coal ash content A c - clean coal ash content At - tailings ash content c,, c 2 - surfactant concentrations in the coal /water suspension and in the filtrate C 1 , C 2 - oil droplet concentrations in the solutions prior to and after abstraction D - amount of oil attached per unit weight of coal, or a constant dV/dt - filtration rate in volume dV - a change in the flow volume from the cake dS - a change in cake saturation F - flocculation percentatge of coal particles in a flocculation test g - gravity constant G.I. - Gaudin Intex (selectivity) A G - free energy change of a system k - permeability of a filter bed k' - a constant characteristic of a filter bed and the fluid properties k" - a numerical constant of a filter bed k s - cake permeability at saturation L - thickness of a filter bed m - weight of adsorbent xix N - number of moles of surfactant adsorbed per unit weight of adsorbent p - applied pressure on a filter cake layer p c - a capillary pressure Ap - pressure drop across a filter bed R - filter cake specific resistance R' - cake resistance R c - a dimensionless capillary number, the ratio of a desaturating force to a capillary force which retains a liquid in a porous media r - a capillary radius S p - specific surface area of a packed filter bed of particles SK - final saturation of the filter cake T - transmittance of a solution V - volume of fluid flowing in time t X - an average value of a group of test measurements X ; - a single test value a - a centrifugal acceleration number T) - viscosity of a fluid s - a fraction of a filter bed volume not occupied by solid material (or voidage) pw - density of a wetting fluid y - liquid surface tension 0 - surface contact angle y o w - oil/water interfacial tension xix A C K N O W L E D G M E N T XX First of all, I would like to express my sincere gratitude to my academic supervisor, Dr. J. S. Laskowski, for his great help, encouragement, and guidance throughout the course of the project. M y gratitude extends to the members of my supervisory committee, Dr. G . W. Poling, Dr. R. W. Lawrence, Dr. B . Klein, Dr. M . Veiga (Dept. of Mining and Mineral Process Engineering) and Dr. B . Bo wen (Chemical Engineering), for their helpful criticism and suggestions concerning this research. I would also like to thank the technical staff in the Department of Mining and Mineral Process Engineering (UBC), especially Mrs. Sally Finora, Mr . Frank Schmidiger and Mr. Pius Lo , for their kind assistance during this work. I am grateful to Mr . Mike Iliffe (Senior Process Engineer, Bullmoose Operating Corporation of Teck Corporation), Miss Kerry Iliffe (English Literature Editor), and Dan Desrosiers (Process Engineer, Bullmoose Operating Corporation) for their valuable help in editing and proof reading of the thesis. M y sincere thanks also go to my wife, L in Lin , for her understanding, encouragement and support. XX 1 C H A P T E R 1 INTRODUCTION Coal is a crucial source of energy for many industries. However, in order to be rendered useful as a fuel or as a feed to make coke, coal must be cleaned to separate mineral matter impurities from carbonaceous organic matter. In present coal preparation plants, gravity separation processes are used for cleaning coarse and intermediate size coal fractions. Froth flotation is employed to beneficiate minus 0.5 mm fines. However, recent environmental restrictions require deep cleaning of coal to further reduce the sulphur and ash contents. The deep cleaning is only possible after fine grinding which produces ultrafine coal (below 40 um) in order to achieve better liberation of coal organic matter from minerals. Since the conventional froth flotation process is inefficient in treating such ultrafine coal particles, new separation techniques, such as selective flocculation, agglomerate flotation and selective oil agglomeration, are being tested. Although these processes have not yet been commercialized by the coal industry, these technologies will play a very important role in the future. More selective and low cost reagents will be needed for the new technologies to be successful. In addition, the dewatering of such ultrafine coal products presents a challenging problem that must be addressed. The processes studied in this research include a selective hydrophobic flocculation process in which hydrophobic latices are used, and a selective oil agglomeration process in which emulsified oils are employed. The subsequent dewatering of the ultrafine clean coal products from these beneficiation processes is also investigated. Selective hydrophobic flocculation is similar to selective oil agglomeration where the oil is replaced by another hydrophobic agglomerant. From this point of view, the process should perhaps be termed hydrophobic agglomeration (Laskowski et al. 1995). 2 Hydrophobic Agglomeration of Ultrafine Coal Hydrophobic selective flocculation (hydrophobic agglomeration) has recently been developed as a new beneficiation process to recover very fine coal particles from a slurry containing both coal and mineral matter. In this process, coal particles are agglomerated by using a hydrophobic latex instead of oil as a bridging compound. After agglomeration, the mineral impurities remain dispersed in the aqueous solution. Consequently, the coal agglomerates can be readily separated from the mineral matter either by sedimentation or by screening. Water-soluble flocculants (e.g. polyacrylamide) were shown to flocculate all particles irrespective of their surface properties. Partially hydrophobic flocculants such as polyethylene oxide (PEO) and F1029-D have been found to flocculate selectively to a certain extent inherently hydrophobic solids. However, the selectivity of P E O in the beneficiation of fine coal particles was found to be insufficient to separate coal from other mineral impurities (Simpson 1990, Palmes & Laskowski 1993). A polystyrene latex, which is the product of the emulsion polymerization of appropriate monomers carried out in the presence of appropriate emulsifiers, was tested by Lyadov et al. (1979) and Littlefair & Lowe (1985) and was found to selectively agglomerate coal particles. However, these tests have never been extended to include various coals and missed entirely a very important factor, the effect of coal surface properties on the agglomeration process. In addition, their test results indicated that the polystyrene latex was still not an ideal beneficiation reagent regarding selectivity. A totally hydrophobic latex, F R - 7 A , was investigated by Attia et al. (1985) and considered a good hydrophobic reagent with a high selectivity in the beneficiation of fine coal. However, the effect of coal surface wettability on the hydrophobic agglomeration was not studied by Attia et al. The F R - 7 A latex is no longer produced by Calgon Corporation and its chemical composition is unknown. In 3 order to continue the tests, a new hydrophobic latex of known composition was needed. The U B C - 1 hydrophobic latex, recently synthesized in the Department of Mining and Mineral Process Engineering, The University of British Columbia, was tested and compared with other available selective flocculants in this work. Selective oil agglomeration is another method that can be used to separate very fine coal from gangue particles. Oil is used in this process as a bridging liquid to selectively agglomerate fine coal particles while leaving hydrophilic mineral particles in a dispersed state. The first oil agglomeration process, referred to as the Bulk Oil Trent Process (Perrott & Kinney 1921) was developed during World War I to clean and recover fine coal. A slurry of finely ground coal mixed with oil at a dosage of about 30% of the weight of the raw coal was agitated for up to 15 minutes. The coal particles formed small aggregates which could be easily separated by screening from the aqueous suspension of mineral particles. The Convertol Process (Brisse & McMorris 1958) reduced the agitation time to 30 seconds by using a high shear mill, and produced stronger agglomerates. However, due to the high oil consumption and high cost of the process, it was not commercialized. More recently, a so-called spherical agglomeration process was developed by the National Research Council of Canada. Densified spherical agglomerates could be formed when 5 - 15% oil was added under appropriate hydrodynamic conditions to an aqueous suspension (Capes et al. 1982). Although oil dosage was reduced by half, the process was still encumbered by a fairly high consumption. Oi l levels of 10% or more by dry weight of the coal were needed, especially when dealing with fine coal containing large proportions of minus 40 um particles (Capes & Germain 1989). The high cost of the oil used at the 10 wt% or greater levels was a continuing impediment to the commercial acceptance of oil agglomeration. Thus, research work by the N R C since 1982 has emphasized agglomeration at decreasing oil levels with the production of much smaller microagglomerates (Capes, 1991). Lately, the Agloflot Process (Ignasiak et al. 1990, U S 4 Department o f Ene rgy 1993), developed by the A l b e r t a Research C o u n c i l and P rax i s Engineers Incorporat ion, reduced the o i l consumpt ion to about 1 - 3%. H o w e v e r , a f lotat ion step had to be added as a means o f separating the sma l l hydrophob ic agglomerates f rom hydroph i l i c gangue, w h i c h made the process more expensive. Other efforts have since been made to reduce o i l consumption. The Otisca T Process, developed by K e l l e r & B u r r y (1990), uses pentane (bo i l ing point o f 36 °C) as an agglomerant and then recovers it by d is t i l la t ion to recycle back to the process. The amount o f the agglomerant used is approximately 5 0 % o f coal particle weight. The associated capi ta l costs for this process are quite h igh and the potential hazards o f us ing vola t i le hydrocarbons have further hindered commerc ia l applications o f this technology. Apparen t ly , to reduce o i l consumpt ion w h i l e mainta in ing h igh select ivi ty and h i g h coa l recovery at a l o w process cost is a c r i t ica l issue i n the appl ica t ion o f o i l agglomerat ion by the coa l industry. Th i s requires new reagents and advanced processes. The reagents shou ld be: (i) able to w o r k at m u c h reduced consumptions, (i i) characterized by h i g h select ivi ty , and ( i i i ) produce strong agglomerates that can be recovered by screening. Enhancement of Filtration/Dewatering of Fine Coal Fi l t r a t ion plays an important role i n the dewatering o f fine coa l , and the degree o f d i f f icu l ty encountered i n f i l t rat ion increases w i t h increasing surface area o f the processed particles. W h e n the f inely ground coal is cleaned by a wet process such as selective f loccu la t ion or o i l agglomerat ion, a great deal o f water remains i n the products. The reduct ion o f the moisture content o f ultrafine coal products has been studied by many researchers w o r k i n g i n this f ie ld (V icke r s 1982, Ber l l inge r & A d a m s 1984, and Pa rekh 1987). The effect o f part icle size on fi l trat ion was reported i n literature (Gray 1958, Geer et a l . 1959, and O f o r i et a l . 1989). Howeve r , most o f the research reports on f i l t ra t ion deal 5 with a -500 um coal. In contrast, this work is mainly concerned with -45 um ultrafine coal and the effect of chemical additives on ultrafine coal beneficiation and filtration. The methods for filtration enhancement can be briefly divided into two groups: i) methods involving equipment modification, and ii) methods relating to process improvement. Chemical additives are applied in the latter to modify the solid surface properties and cake structure to enhance the fluid flow through the filter cake. Three types of filtration additives: flocculants, surfactants, and oily hydrocarbons have been well recognized and tested. However, the mechanisms by which they may improve filtration and dewatering are still not well understood. The basic function of a flocculant used as a filtration aid is to aggregate particulate matter and produce larger channels between the flocculated particles (Ruehrwein & Ward 1952, Gray 1958). This improves the permeability of the filter cake and enables rapid removal of the surface water. Surfactants accumulate at the liquid/air interface and reduce filtrate surface tension. As a result, capillary pressure is reduced, allowing filter cake capillaries between particles to drain more readily, thereby reducing cake moisture (Silverblatt & Dahlstrom 1954, Gray 1958). Adsorbed surfactants may also make the coal surface more hydrophobic if they adsorb onto the coal surface (Nicol 1976, Keller 1979). The role of oily hydrocarbons in filtration is to agglomerate particles and to make them more hydrophobic, facilitating the removal of water, thereby increasing the filtration rate and reducing the residual water content in the filter cake (Gray 1958, Nicol 1980). It is also known that fine coal (-500 urn) and ultrafine coal (-45 um) behave differently in dewatering processes, and the reasons for these differences are studied in this thesis. However, the effects of the chemical additives or their combinations in conjunction with coal surface properties on the filtration of the -45 urn ultrafine coal have not yet been studied. This project addresses these specific topics. 6 C H A P T E R 2 R E S E A R C H O B J E C T I V E S The objectives of this dissertation are to systematically study the effects of chemical additives and coal surface properties on the beneficiation and filtration of ultrafine coal. The program includes studying : 1) Fundamental interfacial phenomena: - emulsification of oily hydrocarbons in the presence of various surfactants - the effect of emulsification with surfactants and coal surface properties on the abstraction of oils by coal - the interactions of hydrophobic latices, surfactants and oil-droplets with coal particles - the effect of flocculants and agglomerants on coal surface wettability 2) Flocculation and beneficiation of ultrafine coal: - the effect of flocculant type and coal surface properties on the flocculation and beneficiation of ultrafine coal - the factors influencing beneficiation of ultrafine coal and its filtration - oil agglomeration of ultrafine coal with the use of emulsified oils and dispersants 7 3) Filtration of ultrafine coal: - the effect of flocculant type and coal surface wettability on ultrafine coal filtration - the effect of surfactants, oily hydrocarbons and coal surface properties on coal surface wettability and filtration of ultrafine coal - the effect of hydrocarbons emulsified with different surface active agents on the filtration of ultrafine coal - the effect of pressure and flocculant on the filtration of ultrafine coal - the differences in filtration behavior between -45 um ultrafine and -500 jam fine coal fractions 8 C H A P T E R 3 L I T E R A T U R E R E V I E W 3.1 Surface Properties of Coal Coal is not chemically uniform, instead it is a mixture of combustible metamorphosed plant remains that varies in both physical and chemical composition. The diversity of original plant materials and the degree of metamorphism are the two major reasons for the variety of physical and chemical behavior in coal. There are three main elements contributing to the surface properties of coals (Laskowski, 1982): 1) The hydrocarbon skeleton (related to the rank of coal); 2) The number and type of polar groups (mostly oxygen functional groups, carboxylic and phenolic); and, 3) The content of inorganic impurities. When coal particles are immersed in an aqueous solution, their surfaces acquire electrical charge caused by either the dissociation of surface functional groups or the preferential adsorption of ions. Surface electrical charge plays an important role in fine coal processing. 3.1.1 Coal rank and chemical composition Coalification is the term for the development of the series of substances including peat, lignite, bituminous, and anthracite coals, while rank refers to the degree or stage 9 reached in coalification by a given coal. Many coal classification systems are primarily based on the content of volatile matter which is the loss in weight when heating coal over a short period of time to about 900 °C. The chemical composition of coal changes drastically with the coal's rank. The higher the coal rank, the lower the content of volatile matter. Coal composition changes with its rank. With increase in the rank, the number of functional groups tends to decrease. Consequently, the coal surface reactivity also diminishes. As shown in Figure 3.1.1-1, taken from Whitehurst et al. (1980), aromatic carbon content rapidly increases with the rank from about 60% for lignite coal to over 95% for anthracite. Figure 3.1.1-1 Carbon distribution in coals of varying rank (Whitehurst et al. 1980) 10 3.1.2 Porosity and Surface Area When coal is crushed and finely ground to obtain a good liberation of the organic matter from mineral matter, the total surface area of the coal drastically increases. This large surface area makes the filtration and dewatering processes more difficult and results in a higher moisture content in the final beneficiation products. The pore network constitutes a significant fraction of the total volume of the coal. Some coals are characterized by specific surface areas which can be as large as hundreds of square meters per gram. According to Gan et al. (1972), who studied American coals from lignite to anthracite, the pore sizes in coals can be divided into macropores (300-30,000 A), transitional pores (12-300 A) and micropores (4-12 A). The relative abundance of these pore size types is related to the carbon content of the coal, and hence to coal rank. In the lower-rank coals (carbon content less than 75%), porosity is primarily due to the presence of macropores. In coals having a carbon content in the range from 76 to 84%, about 80% of the total open pore volume is due to the transitional pores and micropores, while micropores predominate in coals of higher carbon content. Coal moisture content depends to a large extent on coal porosity. Moisture contents at high humidities correlate well with coal porosity, but the moisture contents at low relative pressures characterize interactions between the solid and vapours and depend not only on the solid surface area. 11 The definition of coal moisture is simple: it is that amount of water which can be driven off at 100 - 105 °C (Osborne, 1988). However, the coal industry has adopted numerous and very often confusing terms to describe the forms of moisture. Total moisture is the moisture in the coal, as sampled, determined under standardized conditions ( A S T M D3302). The moisture for coal samples collected from a freshly exposed coal seam face with no visible surface moisture is considered to be inherent moisture. When a coal sample is left in a lab, it will lose some of its moisture (this is the air-dry loss ( A S T M ) , or free moisture (ISO)). The moisture remaining in the sample after it has attained equilibrium with the atmosphere to which it has been exposed is the residual moisture. Equilibrium moisture ( A S T M ) and moisture-holding capacity (ISO) are equivalent in principle and represent a coal's moisture when at equilibrium with an atmosphere of 96-97% relative humidity and 30 °C. Equilibrium moisture results provide a reasonable approximation of inherent moisture for bituminous and some subbituminous coals (Luppen et al. 1991). For example, determination of the moisture immediately after filtration will give the total moisture of the filter cake, while moisture determined in a cake placed in a vacuum desiccator at 30 °C over a saturated solution of potassium sulfate (96% relative humidity) for 2-3 days will give an equilibrium moisture content. In all cases the content of moisture is determined after drying at 100-105 °C following A S T M 388 standards. Mraw and Silbernagel (1980) tested three coal samples of various ranks and found that the coal porosity and total moisture content significantly increase with decrease of coal rank. The high moisture content was caused by high porosity and huge porous surface area for the low rank coal. 12 Coal porosity plays an important role in determining the coal moisture content and surface wettability. According to Kaiji et al.'s (1986) experimental data, the pore volumes alone do not give a good correlation with the water-holding capacities. A fairly good linear relation is obtained between the water-holding capacity and the product of oxygen group content and pore surface area. The higher the content of oxygen groups, the higher the coal moisture content. Sorption of water by coal is expected to occur in three steps: monolayer sorption, multilayer sorption and capillary condensation (Lason et al. 1960). Allardice and Evans (1971) found from measurements of water sorption isotherms that the BET monolayer capacity of coal for water correlated well with the number of hydrophilic functional groups and not the surface area, and that the water in the monolayer is linked to the coal by hydrogen bonding. The importance of hydrogen bonds in the monolayer sorption of water was also studied by a heat-of-wetting measurement (Iyengar and Lahiri 1957). It was also reported by Kraus (1955) and Barton, (1972) that the heat of wetting of carbons in water correlated linearly with the surface oxygen content. Many other researchers also reported similar results. Blom et al. (1951) studied the equilibrium moisture contents of coals of various ranks and showed that it increases proportionately with the coal hydroxyl group content of coals. Tashiro et al. (1969) recognized that coals with a small O/C ratio had a lower moisture content. Schafer (1972) showed that the carboxyl groups played a more important role in determining the moisture content and that about 3.7 times as many water molecules were associated with them than with the hydroxyl groups. Allardice and Evans' (1971) results indicated that the water in the monolayer sorbed on coals was bonded to the carboxyl and hydroxyl sites on the coal surface in a 1:1 molar ratio. More recently, by the measurement of total 13 acidity and fioatability of various coals, Fuerstenau et al. (1983) found excellent correlation between oxygen-containing groups and flotation response; a low content of oxygen groups corresponded with a high flotation rate and low moisture content. The much poorer flotation response of the subbituminous coals was due to the high content of phenolic and carboxyl groups. In the study on coal wettability and its correlation with fioatability, Laskowski et al. (1994) showed a good correlation between the equilibrium moisture content and the total acidity. The total acidity was also found to correlate quite well with the advancing contact angle of coal. Almost all the discussed studies pointed out the importance of the oxygen-containing functional groups on the coal surface for sorption of water vapour. 3.1.3 Functional Groups Perhaps the most important factor influencing the surface properties of coal is the content of the functional groups on the coal surface. This affects surface electrical charge and surface hydrophobicity of coal particles. Many researchers have attempted to establish the nature of the surface groups, particularly through specific chemical reactions. Special emphasis has been placed on carbon-oxygen complexes. Several different analytical procedures have been employed to determine the major oxygen containing groups. Generally speaking, most of the oxygen in coals is present as phenolic, carboxyl and hydroxyl groups and the remainder as ether groups (Ihnatowicz 1952 and Blom etal. 1957). 14 It is generally considered that sulphur exists in coal in three different forms: organic sulfur, pyritic sulfur and sulphate sulphur (Brooks 1956 and Meyers 1977). Physical cleaning of coal can only remove a part of the pyritic sulphur content. The percentage of pyrite removal depends on the pyrite particle distribution in coal and its degree of liberation from the coal. Incorporated during coalification, organic sulphur appears as a part of the coal organic matter, and therefore cannot be removed by any physical process. The majority of the organic sulphur in high rank coals is thiophenic; whereas in low rank coals, most is thiolic and sulphidic (Attar 1979). Of the organic sulphur, 18-25% is in the form of aliphatic sulphides in all coals. Sulphate sulphur represents a very small portion of the total sulphur content and occurs in combination either with calcium (gypsum) or iron (melanterite). The amount of nitrogen present in coal ranges from 0.5 to 3%. Bituminous coals generally contain more nitrogen than lignite and anthracite. The nitrogen present in coals appears mainly as pyridine and quinoline derivatives. Even clean coal contains certain impurities. These impurities are inorganic mineral matter which can be subdivided into six groups (Tsai 1982): clays (kaolinite,-illite, and montmorillonite), carbonates, sulfides, oxides, chlorides, and sulfates. Most of these contain surface hydroxyl groups and charged sites. The major functional groups occurring in coal are summarized in Figure 3.1.3-1. Schematic representation of a possible portion of coal surface is shown in Figure 3.1.3-2. 15 OH I R - C O O R SH 0 R - C - R C9 & 0 - 0 R - S - R R - S - S - R • NH 2 Figure 3.1.3-1. Functional groups in coal (Whitehurst 1978) OH CH 3 OH O CH 3 COOH Figure 3.1.3-2. Schematic representation of a possible coal surface The effect of oxidization on the surface functional groups of coal has been widely studied. Investigation of the nature of the chemisorbed oxygen on carbon and graphite indicate that carboxylic, phenolic, and quinone groups may be present. The nature of these groups has been reviewed by Boehm (1966). Thomas and Hughes (1964) pointed out that the oxidation of the edges is usually four to one hundred times faster than that of the basal planes. Henning (1961) stated that the edge atoms are at least 10 1 2 times as reactive as those in the basal plane. When graphite is oxidized, chemisorbed oxygen is considered to occur mainly at the edges. A s 16 these edges constitute the main adsorption centers, oxygen complexes exert a considerable influence on the surface behavior and surface reactions. The initial stages of oxidation have been characterized by the chemisorption of oxygen at readily accessible surface sites and by the formation of acidic functional groups; in particular, -COOH, =CO, and phenolic -OH. If moisture is present, some chemisorbed oxygen will also form peroxide or hydroperoxide complexes. Tronov (1966) postulated a phenol theory of coal oxidation that suggested a reaction sequence, beginning with the formation of phenols and proceeding through carbonyl compounds to acid anhydrides and carboxylic acids. The formation of phenols has been qualitatively confirmed (Yohe 1947). Dry oxidation studies of subbituminus coal by Jensen et al. (1966) have shown that phenolic structures appear well before appreciable amounts of the carboxylic acids are detected. Van Vucht et al. (1955) found an increase in -OH group content during the oxidation of coal. Adams and Pitt (1955) confirmed these results by infrared studies. It may be concluded that the main reaction during oxidation is the formation of phenolic groups from C-H groups on aromatic nuclei. The effect of oxidation on the fioatability of various coals has been established and will be discussed in the following section. 17 3.1.4 Surface Electrical Charge Most substances acquire electrical charge when contacted with a polar medium such as water. Surfaces may become electrically charged by a variety of mechanisms; in the case of coal, two basic mechanisms contribute to the surface charge. The surface may be electrically charged either by the preferential adsorption of ions or by the dissociation of coal surface groups. Because the system as a whole must be electrically neutral, an equal charge of opposite sign to that on the surface is present in the surrounding liquid. In other words, the surface charge is compensated by an equal but opposite charge distributed in the solution; together, these two charged layers are referred to as the electrical double layer. The first model of the electrical double layer was proposed by Helmholtz (1879) and Perrin (1904) as a pair of parallel charged-plates, which was applicable only to metal-electrolyte system of high salt concentrations (greater than 0.1 M). Gouy (1910) and Chapman (1913) independently suggested a diffuse electrical double layer model, in which the concentration of point charges was considered to decrease progressively with distance away from the solid into the solution phase. Stern (1924) modified and combined the Gouy-Chapman diffuse layer model with that of a condenser-like compact Helmholtz model. Grahame divided the compact adsorption layer into the inner and outer layers. As shown in Figure 3.1.4-1, the border of the layer between the Stern and diffuse layers coincides with the outer Helmholtz plane (OHP). The shear plane is assumed to be close to the OHP. When the particle moves in the solution, the Stern layer will move together with the particle. The potential difference 18 between the bulk of solution (arbitrarily assigned a value of zero) and the shear plane is known as the zeta potential, % (electrokinetic potential). As it has been discussed, electrical charge develops on solid particles suspended in water (or any polar liquid). Only at a particular concentration of potential-determining ions in solution, is the surface charge zero; this is so called the point of zero charge (pzc). If the potential-determining ions are H + and OH" (that is pH), which is very often the case, at pH values above the pzc, the surface has a net negative charge, and below the pzc it is positive. IHP OHP Stern layer Gouy-Chapman layer Figure 3.1.4-1 Schematic representation of an electric double layer in the case without specific adsorption Figures 3.1.4-1 shows that an electrostatic potential which has a value of cp„ at the interface, on moving away from the interface decays exponentially to zero in the bulk 19 liquid. The potential decays because the concentration of counter-ions in the diffuse layer is higher than the concentration of co-ions (Figure 3.1.4-2). The measurement of the electrokinetic potential of coal particles may provide some information on coal surface properties and on interactions between the coal particles and reagents used in various processes. This information may be used to explain phenomena observed in beneficiation and filtration processes. e e e © 0 e e a 0 0 Distance (x) Distance (x) Figure 3.1.4-2 Schematic representation of a diffuse electric double layer 20 The zeta potential - pH curves for different coals can be used to determine iso-electric points of these coals. The negative charges result from dissociation of functional polar groups (e.g., carboxylic and phenolic groups) and also from inorganic impurities (e.g., silica). The nature of the functional groups which produce the positively charged sites is not well established. A variety of factors influencing the surface electric properties of coal were reviewed by Laskowski and Parfitt (1989). These factors mainly include coal rank, degree of oxidation, mineral matter content, and chemical additives. By studying the electrokinetic properties of coals varying in ranks, Wen & Sun (1977, 1981) obtained a series of zeta potential versus pH curves, and found that the low rank coal (e.g. lignite) surface was relatively more negatively charged. The same high-volatile bituminous coal, when oxidized at high temperature, was even more negatively charged. The study carried out by Sobieraj and Majka-Myrcha (1980) showed the same trend. A progressive reduction in the i.e.p. with oxidization was clearly demonstrated. In other words, the high rank coal when oxidized behaved like a low rank coal. Laskowski and Parfitt (1989) generalized the zeta potential versus pH diagram (Figure 3.1.4-3). The diagram seems to be reasonable for different coals, and deviations from linearity are attributed to the presence of acidic oxygen functional groups, mostly carboxylic. 21 Figure 3.1.4-3. Generalized zeta potential versus pH diagram for coals of various ranks (Laskowski & Parfitt, 1989) This subject was also studied by Fuerstenau et al. (1983). The correlation between the content of phenolic and carboxylic groups and coal floatability was found to be excellent. It was also reported that the functional groups controlled coal surface wettability; low rank and high oxygen group content corresponded to coals with a high moisture content. To a certain extent, the surface characteristics of coal are similar to those of oxide minerals (Glembotskii 1972). On this basis, Campbell and Sun (1970) studied the electrokinetic properties of bituminous coal, postulating that the coal surface hydrates and then undergoes pH-controlled dissociation to establish a charged surface in a similar manner to oxide minerals. Their results showed that H + and O H - were the potential determining ions for coal. For anthracite and high-volatile coals, the i.e.p. was found to be near p H 5. Some researchers (Campbell & Sun 1970, Wen & Sun 1977) showed that maximum hydrophobicity depends on p H and is maximum at the i.e.p. of the coal. The 22 results of Fuertenau et al. (1983) clearly demonstrate that the maximum flotation response occurs close to the isoelectric point of the coal particles, which also corresponded to the lowest moisture content of the tested coals. In their filtration studies, Mehrotra et al. (1982) found that the highest settling rate of coal particles and the lowest filter cake moisture content occurred at a pH corresponding to the i.e.p. of the coal particles. Jossep and Stretton (1968), however, showed that some coals floated easily at the i.e.p. while others floated best at pH values greater than the i.e.p. Regarding the effect of the mineral matter in coal on the surface electrical charge, generally speaking, the coal surface tends towards a more negative charge with an increase in the mineral matter content (or ash content) because silicates are the most common impurities in coal. In order to modify the surface properties of coal, some chemical additives (e.g. pH regulators, surface active agents and/or dispersants) may be required. These additives may selectively adsorb onto coal particles, altering the surface electric charge. Wen and Sun (1977) studied the effect of amine adsorption on the eletrokinetic potentials of coal, silica and pyrite particles and found that the dodecyl amine at a concentration of 10 3 M could turn all the particle surfaces from originally negatively charged into positively charged over a pH range of 2 to 11. Jowett et al. (1956) used sodium hydrogen phosphate to disperse clays in order to prevent the ultrafine slime particles from coating the coal surface. After the addition of a dispersant, both the coal and clay particles were found to be more negatively charged. These results clearly indicate that the specific adsorption of phosphate anions increases the negative charge of coal and clay particles. It was also found that humic acids could strongly depress fioatability of coal particles since their 23 adsorption make the coal surface strongly negatively charged (Wen & Sun 1981, L i u & Laskowski 1988). 3.2 Interaction of Chemical Reagents with Coal Surface 3.2.1 Effect of Electrolytes and/or Polyelectrolytes (Flocculants) on Stability of Suspensions Solid particles suspended in aqueous solutions remain dispersed if the repulsive forces are larger than the attractive forces. These repulsive forces include electrical and steric solvation forces while the attractive forces include the London van der Waals dispersion force and the recently found hydrophobic interaction forces. A l l can operate together. According to the classic D L V O (Derjaguin-Landau-Vervey-Overbeek) theory (Derjaguin & Landau 1941, Vervey & Overbeek 1948), coagulation depends on the balance between electrostatic repulsion and the Van der Waals attraction forces. The overall effect of the attractive and repulsive forces is best expressed in the form of potential energy versus distance diagrams; i.e., U=U(r). If the potential energy increases as the distance of approach decreases, a corresponding repulsive or colloidally stable situation will result, and vice versa. Figure 3.2.1-1 (1) shows a colloidally stable situation in which curve (a) represents electrical double layer repulsion, curve (b) represents van der Waals attraction, and curve (c) represents the resultant potential energy obtained by adding curves (a) and (b). Because of the strongly repulsive nature of curve (a), it dominates the form of the resultant interparticle repulsion curve (c). In Figure 3.2.1-1 (2) the repulsive term is considerably reduced and as a result, attraction prevails at all 24 distances. This corresponds to a colloidally unstable or coagulating situation. For larger particles of the size encountered in solid-liquid separation, i. e., >1 um, the resultant curve for a colloidally stable system may exhibit a second minimum at some distance from the surface. This secondary minimum (Figure 3.2.1-1 (3)) commonly gives rise to a weak coagulation that is often significant in technical situations. r (1) s t ab le s y s t e m (2) c o a g u l a t i n g s y s t e m ( 3 ) U n s t a b l e la rge par t ic le (>1 um) s h o w i n g a s e c o n d a r y m i n i m u Figure 3.2.1-1. Potential energy curves (a- electrical repulsion, b- van der Waals attraction, c- resultant potential energy) Although the D L V O theory has been generally successful in describing the stability of colloidal particles, "structural" hydration and hydrophobic interactions were not included in the theory. The structural forces are known to be important at close distances and produce either repulsion in liquids that are able to wet the substrate, or attraction when the solid is poorly wetted by the liquid (Churaev & Derjaguin 1985). Hydrophobic structure forces may cause a very strong attraction (Israelachvili & Pashley 1982) which depends on the hydrophobicity of the interacting surfaces (Churaev 1995 and Yoon 1997). According to Churaev and Derjaguin (1985) and Churaev (1995), approximately at contact angle 9 < 20° the hydrophilic repulsion forces contribute 25 significantly to the free energy of interaction, whereas at 0 > 40° the hydrophobic attraction forces dominate. The D L V O theory is applicable to lyophobic colloids when 0 is between 20° and 40°. When Yoon et al. (1996, 1997) measured the surface forces between two mica surfaces and between a bare glass and a bare silica plate, only "short-range" hydrophobic forces were observed with decay lengths ranging from 0.3 to 3.0 nm in an aqueous solution. However, a longer range hydrophobic force between the mica surfaces with a decay length of 5.5 nm was recorded in a dodecylamine solution. In the presence of dodecanol in the dodecylamine solution, an increasing long-range hydrophobic force was measured with a decay length of as large as 9.0 nm. When the surfaces of glass and silica were hydrophobized with octadecyltrichlorosilane (OTS), much longer hydrophobic forces with decay lengths in the range of 2 - 32 nm were observed (Yoon et al. 1997). Their results clearly illustrate that the n o n - D L V O attraction (hydrophobic) forces may be attributed to hydrophobic interaction, and the hydrophobic force is uniquely determined by contact angle. Especially when 9 >90° , the hydrophobic attraction forces increase dramatically. Therefore, for a colloidal system, a generalized or extended D L V O theory equation should include not only the London-van der Waals dispersion force and electrostatic force but also the hydrophobic attraction forces. When the average advancing contact angle of the two interacting (similar or dissimilar) surfaces exceeds about 90°, the hydrophobic force will dominate the interaction (Yoon et al. 1997). In addition to the hydration forces which are not considered in the D L V O theory, another force - a steric repulsive force exists between adsorbed polymer layers when high molecular weight polymers such as proteins, gums and other synthetic polymers are used 26 to stabilize a dispersion. The steric force appears in the presence of some dispersants or over-dosed water-soluble polymers, and will not be discussed here. The effect of the zeta potential of particles on the stability of a suspension has been well documented. It is known that a stable mineral dispersion can be made to coagulate slowly or rapidly either by reducing the magnitude of the surface potential (by adjusting the concentration of potential determining ions, PDI) or by increasing the ionic strength of the electrolytes present in the medium. When the double layer is compressed, the particles cluster together under the action of the attractive forces. In accordance with Schulze-Hardy rule (Hiemenz 1977), the effective coagulants are electrolytes which contain multivalent cations, such as Ca 2 + , A l 3 + , Fe 3 + , Th 4 + , etc. The action of polymeric flocculants (or polyelectrolytes) is different from that of electrolytes. The coagulation process caused by reduction in zeta potential is not very important in polyelectrolyte flocculation. This is demonstrated by the efficiency of neutral polyelectrolytes as flocculants and by the high efficiency exhibited in some systems in which the flocculant carries the same charge as the flocculated solid particles. However, reduction of the zeta-potential by addition of an appropriate electrolyte prior to the use of the flocculant improves its performance. Much work has been done which confirms the bridging mechanism of flocculation. Ruehrwein and Ward (1952) proposed that the principal effect of polyelectrolytes was to form "bridges" from one particle to the next. According to this model, the polymer chain is adsorbed on the particle surface at only a few points, leaving either ends or loops extending into the solution for contacting the other particles, and thus 27 forming molecular bridges between the adjoining particles in the floe. Akers (1975) postulated that for bridging to happen, the following elementary steps must occur: i) dispersion of the polymer in the solution phase, ii) transport of the polymer to the solid surface, iii) adsorption of the polymer on the solid surface, and vi) collision of particles bearing the adsorbed polymer to permit bridge formation. The action of cationic polymers on negatively charged particles can be somewhat similar to coagulation, in that charge neutralization is the predominant mechanism over polymer bridging. A "charge patch" model for a cationic polymer has been proposed by Gregory (1966). In this model, electrostatic interaction occurs between the localized areas adsorbing and unadsorbing the polymer that has a charge of opposite sign to the surface. However, this does not necessarily mean that the bridging mechanism is absent, in fact it is always operative, especially with high molecular weight polymers. 3.2.2 Selective Flocculation and Hydrophobic Agglomeration Selective flocculation aims at recovering valuable minerals from finely ground ore. The process utilizes the differences in the physical-chemical properties of various fine mineral components of the suspension. It is based on the preferential adsorption of an organic polymer on particular minerals to be flocculated, leaving the remainder of the particles in suspension. The application of this separation technology in fine coal processing relies on the differences of surface properties between organic and mineral matter. 28 A s previously discussed, the assessment of the stability of a suspension based on the relation between van der Waals attraction and electrostatic repulsion is the core of the D L V O theory. The hydrophobic interactions and "structural" hydration forces have been recently added to the extended D L V O theory in addition to van der Waals dispersion and electrostatic forces. In a selective coagulation process in which only p H regulators are used to modify the electrical charge of coal and gangue particles, at a certain p H , most of the coal particles may coagulate together to form aggregates and the gangue particles may still remain in the aqueous solution in a dispersed form. This process has been called "selective hydrophobic coagulation" since the hydrophobic interaction energy is the major driving force for coagulation (Honaker and Yoon 1991). It has been suggested (Laskowski 1995) that all aggregation processes initiated by hydrophobic interactions with the use of the chemical additives, such as hydrophobic latices and oily hydrocarbons, be termed "hydrophobic agglomeration". Accordingly, these hydrophobic additives are referred to as hydrophobic agglomerants. Both selective oil agglomeration and selective hydrophobic flocculation of coal result from hydrophobic agglomeration. Selective flocculation of coal from shale has been studied in the past by a number of authors. In late 1960s and early 1970s, non-ionic and partially-ionic polyacrylamide flocculants were tested as selective flocculants to selectively flocculate coal from mineral matter. Unsatisfactory results, however, were reported. Korczagin (1969), Mozgovoi (1969), Blagov (1970) and Hucko (1977) found that a large amount of shale was always entrapped and flocculated in the coal floes. Either high coal recoveries with high ash contents, or low coal recoveries with relatively low ash contents were obtained in their 29 tests. B l a s c h k e (1972) d i d very interesting reverse f loccula t ion tests i n w h i c h the shale component was f locculated by a po lyacry lamide (Gigtar) f locculant and carboxymethylce l lu lose was used as a dispersant. M o r e recently, Barbery & D a u p h i n (1987) and M o u d g i l (1992) also tr ied to use po lyac ry lamide and polyethylene ox ide to select ively flocculate fine coa l . However , these tests clear ly demonstrated l o w select ivi ty for both f locculants . A fundamental study (Ems ley 1981) o n the o r ig in o f non-select iv i ty o f po lyac ry l amide po lymers showed that al though po lyacry lamide polymers exhibi t some select ivi ty towards either coa l or shale surfaces, there is inevi tably some adsorption o f the f locculant onto minerals by virtue o f the powerful hydrogen bond ing o f the amide group. T h i s group is present i n po lyacry lamides and subsequently leads to non-select ive f loccu la t ion o f a l l particles. S ince hydrogen bonding is a major force between the po lyac ry lamide p o l y m e r and hydrated surface i n aqueous suspensions, it has been suggested (Li t t lefa i r and L o w e 1985) that the o n l y method o f ove rcoming the p rob lem o f non-select ivi ty o f adsorpt ion w o u l d be to use po lymers that have no hydrogen bonding capabil i t ies. The use o f to ta l ly hydrophobic po lymers as selective flocculants presents an alternative to convent iona l polyelectrolytes , w i t h the poss ib i l i ty o f h i g h selectivi ty since the major difference o f surface properties between coal and gangue is their hydrophobic i ty . The effect o f the adsorption o f po lyacry lamide , p o l y a c r y l i c a c i d and h y d r o x y p r o p y l cel lulose o n coa l surface wettabi l i ty was investigated by Prad ip and Fuerstenau (1987) through contact angle measurements, H a l l i m o n d tube f lotat ion and adsorpt ion studies. A s their results showed, the adsorption o f po lyac ry lamide reduced the 30 contact angle of the coal surface; the contact angle decreased from original 6 5 ± 2 to 50 degrees at about 100 ppm polymer concentration and remained around 30 degrees after the concentration reached as high as 1600 ppm. The addition of polyacrylic acid also caused similar behavior. The adsorption of hydroxypropyl cellulose reduced the contact angle more significantly. The flotation results indicated a significant depression of the coal flotation due to the adsorption of hydroxypropyl cellulose; the flotation recovery was reduced from 85% to 25% in the presence of hydroxypropyl cellulose. They postulated that adsorption of the polyelectrolyes on the coal particles resulted from the hydrophobic interaction between the polymers and the coal surface. Moudgil (1983) suggested the same adsorption mechanism. In this type of bonding, most of the hydrophilic groups on the molecule should face the bulk solution, thus making the coal surface hydrophilic. P E O was also used as a partially hydrophobic flocculant by Rubio and Kitchener (1977). They showed that P E O could flocculate inherently hydrophobic solids. Gochin et al. (1985) confirmed this result in experiments in which anthracite was shown to adsorb large amounts of P E O and result in the anthracite being easily flocculated. However, when oxidized, the same coal sample was not flocculated with P E O under any of the test conditions. The tests with totally hydrophobic flocculants in coal beneficiation commenced in late 1970's, and this turned out to be one of the most promising options in fine coal beneficiation in recent years. Lyadov et al. (1979) first used a totally hydrophobic polystyrene latex in the selective flocculation of coal slurries. Subsequently, Littlefair and Lowe (1985) tested polystyrene latex to selectively flocculate coal particles from a 1:1 31 mixture of coal and shale with an ash content of 47%. Several dispersants such as calgon, sodium silicate, sodium polystyrene sulfonate and polystyrene maleic anhydride sulfonate were included. O f these, however, only calgon was found satisfactory for the selective dispersion of shale. The particle size of the coal sample used was 95 wt% -45 um. At a polystyrene latex dosage of 3700 g/t, a flocculated clean coal with a 24% ash content was produced at a 97% combustible recovery by one-stage flocculation. It was considered that there was no hydrogen bonding between the hydrophobic polymer and coal surface which meant that the polymer had only an affinity for the hydrophobic surface, and entrapment of gangue in floes was reduced to a minimum. Entrapment of gangue particles in floes was found to be a major problem in the separation of the floes from gangue particles. Freeing entrapped particles by a second-stage flocculation was suggested and tested. The results showed that a clean coal with a 14% ash content at a 87% combustible recovery was obtained through a two-stage flocculation. Although the results with polystyrene latex were much better from those in which P E O was used, the selectivity was not satisfactory. Attia et al. (1987, 1991) employed a different type of totally hydrophobic latex, namely FR-7A. High selectivity was achieved with the use of F R - 7 A on Upper Freeport and Pittsburgh No. 8 coal samples. Under optimized conditions with a pre-cleaned coal sample, the ash content was reduced from 14.8% to 5.6% at a 95% combustible recovery when the dosage of F R - 7 A was 200 g/t. The effects of process parameters such as slurry p H , solids content, polymer concentration and flocculation time were also investigated. A high recovery was found over a neutral p H range from 6 to 8 and at a shear rate of 170 s_1. 32 However, Attia tested only metallurgical coal. The effect of coal type with different surface properties on selective flocculation was not studied. The process of separation of coal floes from dispersed gangue particles was not investigated either. At present, F R - 7 A is no longer produced and its chemical composition has not been disclosed. Oi l agglomeration of coal is another promising process for separating fine coal particles from mineral matter. As with selective flocculation, this process also relies on differences in the surface properties (hydrophobicity) of coal and mineral impurities. Perhaps the first oil agglomeration application in coal preparation was the Trent process (Perrott and Kinney, 1921). In this process, a mixing vessel was used to agitate a mixture of finely ground coal and water to oil at about 25 - 30% by mass of the coal contained in the slurry. After 10-15 minutes of agitation, agglomeration had proceeded to the extent that the coal agglomerates could be readily separated from the dispersed shale by a screening unit. In 1954, Lemke introduced the Convertol process, in which a high-speed mill was used to create the same but more effective agitation. The agitation time was drastically reduced to 1-2 minutes. The coal agglomerates were removed from the slurry by centrifugation. However, the high oil consumption and high cost of the process made the oil agglomeration technology unfeasible. In 1970s, interest in oil agglomeration was renewed because of the increasing mechanization of mining and the increasing amount of fines in the raw coal, as well as high coal prices and low oil cost. A modified oil agglomeration process, developed by the National Research Council of Canada, resulted from interest in densified spherical agglomerates that formed when a suitable amount of bridging liquid was added under appropriate agitation to an aqueous suspension (Capes et al. 1983). The agglomerated 33 product was easily separated on a screen from the slurry. In this process, two stages of agglomeration were employed. A high shear agitation dispersed the oil phase and promoted contact between the oil droplets and coal particles. Small agglomerates were produced as a result of contact between the oil-coated particles under intensive agitation. The rate of growth of these initially formed agglomerates depended on the amount of oil added, the degree and type of agitation, and the size distribution of the coal. In the second stage, a low shear agitation was applied to enlarge the initially formed small agglomerates. With adequate oil and ideal formation conditions, the unconsolidated floes grew to progressively larger, densified spheres, eventually turning into pasty lumps in which the solids were essentially dispersed in the bridging liquid. Although oil consumption was already reduced by about 50% compared to previous processes, the process still suffered from a fairly high consumption of the agglomerating oil. Oi l levels of 10%o or more of the dry weight of the coal were often used, especially when dealing with fine coal containing large proportions of minus 40 um particles (Capes & Germain 1989). The agglomeration technologies developed by the N R C were implemented at two sites in Pennsylvania. One of the sites, the Florence Mining Co. cleaning plant, operated an agglmeration/screening operation for over two years (Capes 1991). However, the high oil consumption made this process uneconomical. Thus, research work carried out by the National Research Council since 1982 has emphasized agglomeration at ever-decreasing oil levels with the production of much smaller microagglomerates (Capes 1991). A new technology, a slightly modified version of the oil agglomeration process, in which the agglomeration step is followed by flotation, was developed by the Alberta Research 34 Council (Ignasiak et al. 1990) and Praxis Engineers Inc. (U.S. Department of Energy 1993). This low oil agglomeration with flotation recovery process was more selective and produced a product with low impurities. The weaker microagglomerates produced at lower oil levels probably contained only the more hydrophobic and lower ash particles to give lower impurity levels at correspondingly-reduced yield (Capes 1991). The agglomeration flotation of fine coal can be dramatically improved when flotation columns are used (Al-Taweel and Kasireddy 1989, Hirajima et al. 1990). Although the oil consumption might be reduced in the agglomeration / flotation process to about 1 %, a flotation step was added as a means of separation of the hydrophobic agglomerates from hydrophilic gangue. Therefore, the process was still expensive. In order to reduce oil consumption, recycling of the agglomerant liquid has been tested. Keller et al. (1990) developed the Otisca T process, in which pentane (boiling point of 36 °C) was used as an agglomerant, and then recovered by distillation and recycled to the process. The amount of the agglomerant used in this process was about 50%. However, the associated capital costs were quite high and the potential hazards of using volatile hydrocarbons hindered further commercial application of this technology. Fundamental studies (Capes and Germain 1989) pointed out that three major factors controlled oil agglomeration: i) the surface wettability of the solids; ii) oil dosage and type; and iii) the intensity of agitation. Recent studies (Drzymala et al. 1986, Allen et al. 1990) also reported that the presence of air had a beneficial effect on the oil agglomeration of hydrophobic particles. The wetting and bridging between individual particles by a bridging liquid in the agglomeration process was investigated by Jacques et al. (1979). After a thermodynamic 35 analysis of the two possible configurations of initial attachment between the particles and oil droplets, it was concluded that the first configuration, in which the individual particles are totally wetted or covered by the bridging liquid once the bridging liquid droplets attach to the particles, is thermodynamically stable when 0 < 9 < TT/2 (9 is the three phase contact angle measured through the bridging liquid) and n -*• oo (n is the ratio of the bridging liquid droplet size to the solid particle size). Essentially, this means that the more hydrophobic the solid and the larger the amount of the bridging liquid, the more likely the total wetting process will occur. Compared to the first configuration, the second configuration, in which the two individual particles are bridged by a single droplet, is thermodynamically more likely. It was shown that for particles preferentially wetted by the bridging liquid, the bridging is only possible when 0 < 9 < n/2, and that a definite limit on the range of 9 exists for each value of n. The most thermodynamically stable bridges were found to form for a small contact angles (9) and for large values of n. In other words, the more hydrophobic the solid particles, and the larger the bridging liquid volume, the more likely is the formation of the bridges. In the case of hydrophilic particles suspended in an organic phase, water will similarly bridge the hydrophilic particles. Since coal is heterogeneous and is composed of a mosaic of strong and weak hydrophobic sites and hydrophilic sites, it is inevitable that some water will always be trapped in the agglomerates during the oil agglomeration process. The oil dosage level is the most important factor in oil agglomeration. At low oil levels, only pendular bridges form between particles yielding a loose and unconsolidated floe structure. Discrete lens-shaped oil rings and points-connected oil bridges are formed 36 at the points of contact of the particles. At high oil levels, the oil rings and discrete oil bridges begin to coalesce and form a funicular bridge - a continuous network. Microagglomerates will form at this stage. Interparticular space will be either fully or partially filled with the bridging oil. When the oil filling the space between the particles reaches a maximum value, the microagglomerates will combine into pasty lumps or spherical agglomerates. At the same time, the formed agglomerates are consolidated. Different oil agglomerating mechanisms have been postulated. According to Capes (1982) and Szymocha & Ignasiak (1989), at the beginning of the agglomeration process, oil droplets first attach onto individual coal particles. During this period no growth of agglomerates is observed. Intensive mixing of the slurry results in dispersing the bridging oil into fine droplets which are deposited and spread to form a 'film' on the individual hydrophobic particles. In the quick growth period, contact between the oil-coated particles occurs and pendular floes are formed. However, a thermodynamic analysis of oil spreading on solid particles (Laskowski, 1991) indicates that the spreading of oil into a thin 'film' at the solid/water interface is thermodynamically always unlikely. In line with this conclusion, Klassen (1963) shows only oil droplets or some oil patches appearing on a hydrophobic surface in water. Keller and Burry (1987) also disagreed with the mechanism proposed by Capes (1982). The coal surface is a patchwork assembly which consists of areas of various organic matter, mineral matter and pores. It can display various contact angles with water. The spreading of oil proceeds to the limit of an area in which water is hydrogen-bonded to the particle surface. While hydrocarbon liquid bridges are formed between the hydrophobic parts of coal particles, water attaches to the patches on the coal surface at points where exposed mineral matter imbedded in the coal 37 intersects, or fills the pore in the coal surface. In a study concerned with the elimination of pyrite from coal in oil agglomeration by using starch, a similar agglomeration mechanism was proposed by Good et al. (1994). Relatively speaking, coal particles are more hydrophobic than pyrite. Although oil droplets preferentially attach to the coal surface during oil agglomeration, they also attach to pyrite particles. When the partly oil-covered particles collide with each other, the oil that is on them merges, and the coal particles form agglomerates in which there is a continuum of oil, and water droplets exist between the coal particles. However, the oil between the pyrite particles only forms droplets, whereas water becomes a continuous phase in the pyrite agglomerate. Therefore, the configuration formed by the oil-water-coal system leads to stable macroscopic agglomerates, but the configuration of the oil-water-pyrite system is relatively unstable. As agitation continues, the aggregates that contain pyrite tend to break down (instead of growing) and to release pyrite particles into the aqueous phase, where they remain suspended. Other hydrophilic species such as clay, silica, shale, limestone, etc., behave similarly to pyrite. The type of oil used is as important as its concentration in the agglomeration process. The major differences among all types of oils are their densities, viscosities and chemical compositions. It has been found, however, that of those characteristics, the density and viscosity play more decisive roles. The effects of oil density and viscosity on agglomeration are due primarily to the hydrodynamics of the system, rather than to the chemical nature of the bridging liquids (Capes and Germain 1989). During their experimental work on the Convertol process, Sun & McMorris (1959) concluded that oils with medium density were required to obtain satisfactory recovery levels for coal fines. 38 Low-density oils were considered to have insufficient viscosity to "pull" the coal particles together into strong agglomerates. On the other hand, high-viscosity oils were not dispersed sufficiently well in the slurry to enable particles to wet and cause agglomeration. Germain (1975), in his experiments with three paraffmic oils of increasing density, showed that poor coal recovery was found with the most dense oil when the mixing time (1 min) was not sufficiently long. However, with a longer agitation time, this oil yielded an excellent coal recovery. Under sufficient mixing conditions, the viscous oil also produced relatively larger agglomerates, which in turn contained less moisture and had a lower ash content. However, Capes' results (1976, 1977) showed a general trend in which oils with lower densities (below 0.9 g/cm3) generally exhibited better ash rejection and lower coal recoveries, while oils with higher densities (greater than 0.9 g/cm3) showed more erratic behavior with substantially higher ash products and relatively higher coal recoveries when sufficient mixing was used. Capes explained that the resultant high ash products were due to the viscous heavier oils used in the tests. These included complex mixtures such as petroleum resins, tars, bunker fuels, heavier crudes, and coal tar as well as its derivatives, containing higher levels of functional groups. These polar functional groups are apparently able to attach to the ash-forming mineral particles and make them report to the agglomerate products. The interfacial tension for aliphatic hydrocarbons and water is about 50 erg c m 2 , while any content of aromatics and especially hydrocarbons with polar groups will lower this value appreciably down to 15-16 erg cm - 2 or smaller (Laskowski 1991). The aromatic hydrocarbon molecules that contain some polar groups interact with water and mineral impurities through hydrogen bonding. Various types of interactions displayed by 39 aromatic, naphtenic and oxygen-, sulfur- or nitrogen- containing hydrocarbons permit them to adsorb onto various solids, which also contributes to the higher ash content of the agglomerates. However, the heavier fractions from crude oil distillation are shown to be better agglomerants for low rank and/or oxidized coals. The effect of agitation time on agglomeration has already been mentioned. The importance of mixing time increases as oil density and viscosity increases. The mixing time needed to form adequate agglomerates decreases as agitation is intensified (Capes 1982). The purpose of agitation is to disperse the bridging oil into small oil droplets, deposit them to the coal particles, and create collision of oil-coated particles. Bensley and Nicol et al. (1977) showed that emulsification of the oil improves oil agglomeration. They studied the effect of prior oil emulsification on process parameters such as inversion time, product ash, and recovery of organic material. A l l of the oil emulsions were prepared by mixing water and oil in a high speed blender for 3 minutes, or in a mixer for 60 minutes. The emulsions reduced the inversion time by about 50% and increased the coal combustible recovery by 1 - 5%. However, no reduction of oil consumption was reported. The results suggested that the emulsions had no effect on the product ash content, either. This may result from the fact that surfactants were not utilized in the emulsification. The mechanically produced emulsions were found to be too unstable, and the oil droplet size rapidly increased even a few seconds after emulsification, thus only an approximate size could be given in the experiments. Coal wettability affects both flotation and oil agglomeration in a similar way. Lower rank coals such as subbituminous and lignite are distinguished by their greater oxygen content and the hydrophilic nature of their surfaces relative to those of 40 bituminous coals. Capes et al. (1976) found that light oils which successfully agglomerate bituminous coals were unable to agglomerate lower rank coals, while heavier oils which contain hydrophilic functional groups such as nitrogen, oxygen and sulfur polar groups, could be used to agglomerate low rank coals as well. Coke oven tars, pitches and petroleum crudes were also used in these agglomeration tests. Apparently the polar groups of these complex oils play an important role: they are able to adsorb on the relatively hydrophilic surfaces of low rank coals and sufficiently well to form agglomerates. The importance of the aromatic content of oil for agglomerating low rank coals was emphasized by Blaschke (1990). Her results show that while metallurgical coals are well agglomerated by diesel oil, subbituminous coals can be cleaned by oil agglomeration i f the used oil contains aromatic hydrocarbons. In addition, as found by Pawlak et al. (1985), thermal and subbituminous coals were well agglomerated when oil containing bitumen or heavy refinery residues were utilized. Labuschagne (1987) studied the influence of oil composition and pH on selective agglomeration of coals with different wettability. As shown in his results, benzene and cyclohexane, which perfectly agglomerated the hydrophobic coals, could not agglomerate the more hydrophilic coals. The hydrophilic coals, however, were well agglomerated when a small amount of alcohol was used with these bridging liquids. This improvement in agglomeration was attributed to the interactions between the hydroxyl group of the alcohol and coal surface sites. Brown et al. (1958) studied the spreading of oil on a wet coal surface, hoping to find conditions under which the spreading coefficient or the work of spreading was approaching zero. (Oil would spread spontaneously on a wet coal surface under such 41 conditions.) Actually, the purpose of their study was to find reagents that would improve the hydrophobicity of coal particles in the flotation process. It is generally considered that the easier the oil spreads on the wet coal surface, the higher the recovery. However, Brown found that the best flotation oils were those which spread least readily, though no explanation was offered. Moxon's study (1987) showed that the maximum flotation yield was obtained with a hydrocarbon chain length corresponding to dodecane. The low yield obtained with shorter chain oils was considered to be retarded by the penetration of the oils into the coal pores, thereby reducing the surface concentration of oil. On the other hand, longer chain oils also produce low yields due to the greater viscosity of these materials which reduced the rate of spreading on the coal surface. The viscosity effect could be compensated for by providing longer conditioning time. It is noteworthy that the effect of various oils on flotation of coals corresponded very well to the results of oil agglomeration. Surface active agents, which are usually used as flotation frothers, were also tested by Brown et al. (1958) in his contact angle measurements. It was found that these frothers could reduce the contact angle of oil on a coal surface in water and decrease the oil/water interfacial tension, thereby making the spreading of oil on the coal easier. The spreading coefficient of paraffin on coal surface was significantly increased due to the addition of m-cresol to the oil. Laskowski's experiments (1986) show that promoters reduce the kerosene/water interfacial tension from 50 mJ/m 2 down to below 1 mJ/m 2 with the 5% addition of Dowell M-210 to kerosene. This leads to improved spreading of kerosene onto the coal surface. 42 3.2.3 Interaction of Surfactants with Solid Surface and Adsorption Mechanisms 3.2.3.1 Interaction of Surfactants with Solid Surface Surfactants contain both a polar hydrophilic group and a nonpolar hydrophobic group. The nonpolar group is usually represented by a hydrocarbon chain. It has no permanent dipole and constitutes the hydrophobic part of the amphipatic molecule. Since the two parts of the surfactant molecule have an affinity for different phases, the surfactant molecules orient themselves at the air/water interface with the polar group incorporated into the aqueous phase and with the hydrophobic radical oriented away from the water. Accumulation of the surfactant molecules at the interface brings about a reduction in the interfacial tension between the two phases. V HYDROPHILIC SURFACE Figure 3.2.3.1-1. Model for adsorption of surfactant molecules onto hydrophobic and hydrophilic solid surfaces (from Clunie and Ingram 1983) 43 When solid particles are present in the surfactant solution, surfactant molecules may adsorb at the solid/liquid interface. Adsorption of the surfactant on solids from solution is important in many practical situations. Perhaps modification of the solid surface is of primary concern (Figure 3.2.3.1-1). The adsorption of surfactant at a solid/liquid interface is generally considered to be controlled by three factors, which are related to the materials in question, namely: i) the chemical nature of the species being adsorbed, including the nature of the polar group (anionic, cationic, nonionic, etc.) and the nature of the hydrocarbon chain (length and degree of branching, etc); ii) the nature of the solid surface onto which the surfactant is being adsorbed (highly charged, nonpolar, etc); and iii) the nature of the liquid environment (in water the pH, ionic strength, temperature, additives, etc). The adsorption of a surfactant molecule onto a solid surface can be affected significantly by relatively small changes in the characteristics of the system. Ionic surfactants are often employed as collectors in mineral flotation to render a mineral surface hydrophobic. Such collectors are not required in coal flotation. However, the long chain surfactants, such as alkyl sulfates, alkyl amines and nonyphenols, are used throughout this study. These anionic, cationic and non-ionic surfactants are utilized to modify the coal particle surface, to improve the dispersion of oil in aqueous solution, and to enhance the filtration and dewatering processes. 44 3.2.3.2 Adsorption Mechanisms It is common to categorize the mechanisms of adsorption of flotation collectors at the mineral/water interface into three broad classes (Shergold, 1984): i) Coulombic attraction between the ionized collector species and surface bearing an opposite electrical charge followed by the formation of hydrophobic associations (hemimicelles), e.g. adsorption of alkyl sulfates and sulfonates on oxides. ii) Chemical interaction between the collector and sites in the mineral surface resulting in the formation of a chemisorbed layer or a new phase at the mineral/water interface, e.g. adsorption of fatty acids by salt-type minerals (calcite, fluorite, etc.). iii) Electrochemical interaction between semiconductor minerals and a collector that readily oxidizes. This applies to the adsorption of thio-collectors by sulfide minerals. To make this list complete, a fourth group can be added to these three which includes the attachment of droplets of water-insoluble oily collectors to mineral surfaces through selective wetting, as in emulsion flotation of coal with kerosene or diesel oil, coal oil agglomeration, etc. Whereas electrostatic interactions play a very important role in the adsorption of ionic surfactants, the chain length is important too. This is illustrated by the example of adsorption of sodium alkyl sulfonates onto positively charged alumina (pH 7.2), taken from Wakamatsu and Fuerstenau (1973). Since hemimicellisation results from the association of the hydrocarbon chains, it should take place at much lower concentrations when the length of the surfactant hydrocarbon chain is increased. Figure 3.2.3.2-1 confirms such a conclusion. Relatively short chain octyl sulfonate ions adsorb only 45 through electrostatic interactions onto a positively charged solid surface. For longer chain alkyl sulfonates (e.g. C 1 2 to C l 8 ) , at a certain critical concentration there is a sudden change in the shape of the adsorption isotherm which becomes very steep. This indicates hemimicellisation. Many flotation studies confirm this type of adsorption behavior of ionic flotation collectors (Iwasaki, 1960; Somaundaran and Agar, 1964). Wen and Sun (1977 & 1981), Brooks (1956), Gala (1982) and Keller (1979) studied similar interactions while working with coal and cationic surfactants. 10 Alumina pH 7.2, Ionic strength 2x10(-3) mol/l C10 C8 -13 10 -5 -4 -3 -2 10 10 10 10 10 Equilibrium concentration (mol/l) Figure 3.2.3.2-1 Adsorption isotherms for sodium alkylsufonates with different hydrocarbon chain lengths on alumina at 2x103 mol/l ionic strength (Wakamatsu and Fuerstenau, 1968) The sulfonates and amines used in this project belong to the group of ionic surfactants. It must, however, be borne in mind that while these two are both ionic they 46 are different: sodium alkyl sulfonates are strong electrolytes while primary amines are weak electrolytes. Castro and Laskowski (1986) and Laskowski (1989) showed that the latter might form colloidal systems with quite distinct properties. Nonyl phenols and ethoxylated nonyl phenols belong to the group of nonionic surfactants which, as discussd by Aston et al. (1989), may adsorb onto both hydrophobic surfaces through hydrophobic interactions as well as onto polar surfaces mainly through hydrogen bonding. Rubio and Kitchener's (1976) results indicate that polyethylene oxide (PEO) strongly adsorbs onto hydrophobic silica which has only isolated silanol groups on its surface. Adsorption is favoured i f the regions between such sites are hydrophobic. In other words, their results show that PEO and similar compounds should adsorb onto hydrophobic surfaces that contain some polar groups. Gochin et al.'s data (1985) on the flocculation of anthracite by PEO agree very well with these conclusions. 90 15 A ANTHRACITE 0 0 -6 -5 -4 -3 10 10 10 10 TRITON N-101 CONCENTRATION (mole/I) Figure 3.2.3.2-2 The sessile drop contact angles for a HVA-bituminous coal and an anthracite coal as function of Triton N-101 concentration (from Chander et al. 1987) 47 Coal is a very heterogeneous solid with very different surface sites. As shown by Aston et al. (1989), ethoxylated nonyl phenols can adsorb onto both hydrophobic and hydrophilic surfaces. The orientation of the adsorbed molecules at these surfaces is however very different (Figure 3.2.3.1-1). The usual overall result of the adsorption of such compounds on coal is shown in Figure 3.2.3.2-2 (Chander et al. 1987). As this figure reveals, the adsorption of ethoxylated nonyl phenols onto bituminous coal increases the contact angle at low surfactant concentrations but reduces its value to almost zero at concentrations close to the critical micelle concentration. This explains why such compounds are applied as promoters/emulsifiers at low concentrations, but are used as stabilizers (e.g. in coal water slurries) at high concentrations. 3.3 Filtration and Dewatering of Fine and Ultrafine Coal The purely empirical approach, when filtration or dewatering is studied under varying conditions, has obvious advantages for the solution of practical problems. The results, however, usually lack generality. In contrast, the purely theoretical studies deal only with a mathematical description, which contributes little to the understanding of the mechanism of the process. Very often, it is necessary to conduct experimental filtration tests along with a theoretical analysis of the involved phenomena. The response of a fine coal slurry to filtration and dewatering depends on many factors, including the coal surface properties, particle size distribution, process 48 temperature, solution pH, nature of chemical additives, dosage levels, conditioning time, hydrodynamic conditions, types of dewatering devices, etc. First, an overview of filtration and dewatering theories will be presented. The review of the experimental studies on filtration and dewatering enhanced by the use of flocculants, surfactants, and oily hydrocarbon additives will follow. 3.3.1 Theory of Filtration and Dewatering Filtration processes are conventionally divided into two classes (Perry & Chilton 1973): deep bed filtration and cake filtration. Deep bed filtration is normally preferred when the solids content of the slurry is very low (less than 0.1 %). In this operation, a deep bed of porous media like diatomite or fibrous material, supported on a coarse filter, is used to remove the fine solids from the slurry. The particles to be removed will often be considerably finer than the pores of the filter medium and will penetrate a considerable depth before being captured. These particles can be captured by any of the following mechanisms (Grace 1956): i) direct-sieving action, ii) gravity settling, iii) Brownian diffusion, iv) interception at the solid-liquid interfaces, v) impingement, and vi) electrokinetic forces. In cake filtration, the solid material accumulates on the surface of the medium, so that, after a short initial period, filtration occurs through the bed of deposited solid. This process will continue until the pressure drop across the cake exceeds the maximum permitted by economic or technical considerations, or until the space available is filled. This method of filtration is the most widely employed by industry and is very well suited to the filtration of concentrated suspensions and the recovery of large 49 quantities of solid. The filtration of coal or other mineral slurries employs this method. The most important factor in cake filtration is the permeability or resistance of the filter cake, which may be more or less controlled by altering the particle size distribution of the material, sometimes by adding coarse solid particles to it, and also by altering the state of aggregation of the particles. The techniques in which cake filtration is enhanced by chemical additives (e.g. flocculants, surfactants and oily hydrocarbons) have also been widely used. Compared to filtration, which can be described as a flow of a single fluid (the filtrate) through a porous medium, dewatering by air displacement is a problem of a two phase flow (involving both filtrate and air) through a porous medium. Dewatering is described by Wakeman (1975) as a post-filtration process in which the filtrate is removed from the void space of the filter cake by the application of desaturating forces. The differences between filtration and dewatering are shown in Figure Apdx - 1 for a funnel type of vacuum filter. During the filtration period (Figure Apdx - 1 (a)), a filter cake forms by solid particle deposition, and the suspended particles are moving towards the filter medium. At the end of the filtration period, this relative motion of solid particles ceases and a saturated filter cake is created (Figure Apdx - 1 (b)). During the dewatering period, air is drawn through the filter cake to displace the water. In all filtration operations, however, it is impossible to remove all the water from the cake. A partially saturated state, as shown in Figure Apdx - 1 (c), is finally reached at the end of the dewatering period. 50 3.3.1.1 Theory for Filtration Process under Constant Pressure From a study of the flow of liquids through a bed of sand, Darcy (1856) proposed the empirical relation: A dt L w where (l/A)(dV/dt) is the overall fluid velocity, L is the thickness of the bed, Ap is the pressure drop across the bed, k' is a constant characteristic of the bed and the fluid properties, V is the volume of fluid flowing in time t, and A is the cross-sectional area of the bed. Shortly before Darcy's work was published, Poiseuille (1840) derived the following equation for the flow of liquid through a capillary of a circular cross section: d V _ A p - T r - r 4 dt 8-TJ-L where r is the capillary radius, and n is the viscosity of the fluid. It was soon realized that the viscosity factor appearing in the Poiseuille equation also played an important role in filtration (Dickey 1961). Thus, the modified Darcy equation is written as: 51 where k is defined as the permeability of the bed. The unit of k when (l/A)(dV/dt) is expressed in 1 cmVsec cm 2 for a liquid with viscosity 1 cp and at a pressure drop (Ap) across bed of 1 atm is known as the Darcy. Two other important properties of a packed bed are the filter cake porosity, s, defined as the fraction of the bed volume not occupied by solid material (also known as the fractional voidage or voidage), and the specific surface area, S p , of the packed bed of particles per unit volume of bed. The volume fraction of bed occupied by the solids is 1-s. The two parameters are combined in the cake property called cake specific resistance R. An alternative expression for the fluid volumetric flux through a porous medium, derived by Kozeny and Carman, is: where k" is a numerical constant normally taken as having the value 5 for laminar flow. The cake specific resistance R is therefore expressed as 1 dV Ap (4) A dt k"-S p 2 - (1 - s ) 2 n - L R = (1 - s ) 2 (5) or R = 5 -S„ 2 - ( l - s ) 2 s 3 (6) The cake resistance R' can be calculated from: 52 R - R L (7) The value of R or R' is a property of the particles forming the filter cake. Therefore, i f the support medium resistance is neglected as small in comparison to the cake resistance, from equation (4), the filtration rate dV/dt depends on filter area A , filtration pressure Ap, cake thickness L , filtrate viscosity n, and cake specific resistance R, which is related to the particle size distribution. 3.3.1.2 Theory for Dewatering Filtration of a slurry results in a cake in which the capillaries are filled with filtrate. After the water on the surface of the filter cake disappears, a dewatering step begins. In dewatering, the water in the cake is removed by applying desaturating forces. These forces can be either mechanical, as in dewatering by compression, or hydrodynamic as in the displacement of water by compressed air. The difficulty with which the water can be removed will largely depend on the characteristics of the porous cake. Like the water distribution in a filter cake as stated before, porous systems may also be classified into three categories according to their size: voids, capillaries and "force spaces". These were discussed in detail by Marefold (1937). Vo id systems are easily filled and emptied with foreign substances. Capillary systems offer resistance to filling with foreign substances, and in the end a partition remains 53 behind. The "force space" systems can be filled or emptied only with molecularly dispersed substances and only under great pressure or specific chemical forces. Factors affecting a dewatering process are summarized in Table 3.3.1.2-1 by Wakeman(1975). Table 3.3.1.2-1 Factors affecting dewatering as summarized by Wakeman 0975): Properties of Properties of Interfacial Other the cake the fluid properties factors particle size density interfacial tension temperature particle shape (liquid-liquid) mode of particle viscosity surface tension pressure gradient packing (gas-liquid) dimensions of the rate of displacement cake 3.3.1.2.1 Model for Dewatering Process A flow model for dewatering, used to calculate the kinetics of the process, has been developed by Wakeman (1975). Applying Darcy's law to a layer of the medium, one obtains: 1 dV = k,-(p- pe), ^ A dt -L 54 where p is the applied pressure on the cake layer, p c the capillary pressure, k s the cake permeability at saturation s, and (p-p c) s the pressure drop across the layer of cake. A change in the flow from the cake, dV, causes a change in cake saturation, dS, which is expressed as follows: A - L - e by substituting equation (9) into equation (8), one can obtain: DS=K-(p-P,l-dt ( 1 Q ) * L ' The model shows the relative importance of various parameters during the dewatering operation. It also indicates that the applied pressure, p, must be larger than the capillary pressure, p c , otherwise a desaturation process will not occur. 3.3.1.2.2 Model for the Residual Saturation of the Cake When the filtrate is being displaced by air, not all of the water can be removed from the filter cake at equilibrium. The final saturation of the cake strongly depends on the cake structure, the filtrate and particle properties, and the applied desaturating forces. Brownell and Katz (1947) suggested the following correlation for cakes drained by gravity or centrifugal force: S*, = — . F t - 0 ' 2 6 4 (1 1) 86.3 c 55 where SX is final saturation of the filter cake, and R c is the dimensionless capillary number, the ratio of the desaturating force to the capillary force which retains the liquid in the porous media. The capillary number is given by + w-a + Aplg-L • cos (12) where g is gravity constant, w is the density of the wetting fluid, a is the centrifugal acceleration number, y is the surface tension, and is the contact angle. Equation (11) gives value of S00>1 for very small values of R c (R c <10"8) so it should be applied with care. Brownell and Katz (1947) gave another empirical equation expressing the residual saturation of a coal filter cake as follows: S = D . k-(p-pc) • cos • L -0.264 (13) where D is constant and the other parameters are those mentioned previously. 3.3.1.3 Enhancement of Filtration and Dewatering In the filtration and dewatering of a coal slurry, two important aspects are the rate of filtration, and final moisture content of the filter cake. Equation (4) shows that the filtration rate can be improved by i) increasing the filtration pressure, ii) increasing the cake permeability or decreasing the cake resistance, and 56 iii) decreasing the filtrate viscosity Interfacial properties such as surface tension and contact angle have a strong effect on the amount of filtrate retained in the cake. The basic equation for the capillary pressure in a cylindrical tube was given by Young and Laplace (Adamson 1982). One form of this equation is 2- -cos . . . . Pc = (14) r where r is the capillary radius for very small capillaries. Substituting equation (14) into equation (8) and (13) gives 2 cos i dv k ^ P — A dt -L (15) and S=D' k • , x -1-0.264 ' p 2 A L • cos r • L) (16) From equations (15) and (16), the following general conclusion can be drawn: The dewatering rate increases and the final cake moisture content decreases as (a) driving force (p) increases, (b) cake permeability (k) increases, (c) surface tension (y) decreases, (e) contact angle (6) increases, (f) filtrate viscosity decreases, and (g) cake thickness decreases. 57 In summary, for the enhancement of both the filtration and dewatering processes, the filtration pressure, the dewatering driving forces, and the cake thickness are mechanical factors that can be optimized at a given coal dewatering facility subject to equipment and economic considerations. The reduction of the filtrate viscosity is practically impossible without heating. From the brief analysis of the above theories for filtration and dewatering, it is clear that the following three possibilities, which can improve filtration and dewatering processes by addition of dewatering additives, are open for research: increasing filter cake permeability, increasing hydrophobicity of the particles, and reducing filtrate surface tension. 3.3.2 Filtration Enhanced by Flocculants The application of synthetic polymer flocculants in coal dewatering goes back to the mid-1950s when such materials first became available. From the point of view of capillary theory, Gray (1958) suggested that fine coal (-600 um) dewatering would be improved by flocculation. He used a large number of possible flocculating agents and found that flocculation could increase the mean pore size between particles and the best flocculants were two polymers based on polyacrylamide: Separan 2610 (Dow Chemical Co.) and Aerofloc 552 (American Cyanamid Co.). These increased the initial moisture before vacuum was applied, but gave significant reductions in moisture content under suction. The optimum concentrations of these materials were much lower than those of more traditional and cheap flocculants, such as starch and gelatin. Grey's interest in 58 additives used in dewatering fine coal was not only in flocculants but also in oils and surfactants. Dewatering was found to be improved by each of the above additives. He postulated that the dewatering enhancement could be caused by both flocculation and hydrophobization. Grey's work was, however, confined to one type of coal. Geer et al. (1959) also studied the use of flocculants in dewatering of -600 urn coal. Geer and his co-workers tested three natural slurries and three coal-clay mixtures. Polyacrylamide-based materials and starch in combination with lime were used as flocculants. The effectiveness of synthetic polymeric flocculants at very low dosages was confirmed. It was concluded that the three natural and three synthetic slurries exhibited individual characteristics and individual filtration properties. Moreover, their filtration behaviors were not directly related either to the size distribution or to the ash content. Thus, each slurry had to be evaluated individually for possible benefits of flocculation. However, the results clearly showed that the cake formation rate could be increased by all tested polymeric flocculants, while the moisture content was also increased by the flocculants for relatively fine coal, even with a lower ash content, and decreased for coarser coal even with a higher ash content. The same conclusion was drawn by Meerman (1957), whose investigation was conducted on a full plant scale. Gieseke (1962) tested Aerofloc flocculants at various coal preparation plants and found that the cake moisture and cake formation rate both increased. Osborne (1974) did extensive work on coal-shale slurries and provided ample evidence of the same type of results as Geer et al. (1959). Osborne also made the general observation, common to all flocculants, that both cake 59 formation rate and filter cake moisture were always observed to increase simultaneously. Later, he reported a relationship between ash content and flocculation dosage. Mehrotra et al. (1982) investigated the dewatering characteristics of flocculated coal slurries. A cationic (Cat Floe produced by Calgon Corp.), a nonionic (Separan M G L from Dow Chemical Co.) and an anionic flocculant (Separan MG700 from Dow Chemical Co.) were used. They found that the dosage of the flocculants and the slurry p H had a significant effect on the filter cake moisture, and the lowest filter cake moisture was observed at the p H corresponding to the point of zero charge (pzc) of the coal particles. The usage of the nonionic flocculant reduced the moisture content of the -600 um fine coal from 28% to 18%. It was proposed that non-selective adsorption governed by hydrogen bonding and excessive flocculant dosage were detrimental to dewatering. They also suggested the possibility that flocculation might eliminate the stratification of coarse and fine particles during cake formation, and thus improve the overall dewatering process. Furthermore, they observed that: i) an anionic flocculant was expected to be effective in flocculating the suspension only when the particles were positively charged (i.e., below the pzc); above the pzc, the hydrogen bonding and the increased concentration of cations such as K + could also result in the adsorption of anionic flocculants and flocculation; ii) above the pzc, a cationic flocculant is adsorbed mainly by electrostatic attraction and gives substantial flocculation, but overly-negative charged particles will decrease the extent of flocculation due to the flat conformation of the cationic polymer at the solid/liquid interface; iii) the nonionic flocculant, Separan M G L , 60 is the most effective, and the best dewatering results were achieved at pH value close to the pzc of particles. Mehrotra and Wakeman's conclusion about the lowest cake moisture occurring at a pH corresponding to the pzc of the coal particles was recently confirmed by Groppo et al. (1994), who investigated the effects of pH, cake structure, viscosity and solid content on the high pressure filtration of fine coal. Lewellyn and Wang (1982), in their studies on the effect of molecular weight and degree of anionicity of copolymers prepared from sodium acrylate and acrylamide, found that the optimum dewatering (for cake moisture and cake formation rate) occurred at a relatively low molecular weight (5-8x105 g/mol) and a high acrylate content (70-80%). They also considered looser floes to be detrimental to filtration since they entrapped water and collapsed on the filter medium, thereby blinding it. Perhaps the most disappointing are the results obtained by Nicol (1976), who applied Magnafloc R155 and Alfloc 6731 flocculants, commonly used in Australian coal preparation plants, to aid the filtration of the -500 +76 um coal suspensions. His results showed a decrease of filtration rate and an increase in cake moisture content with the addition of the flocculants. Cheng and his co-workers (1988) examined the effect of various anionic, nonionic, and cationic flocculants and their dosages on the vacuum filtration of -600 um coal fines. Their conclusions can be summarized as follows: i) the best results were obtained with an anionic flocculant with a molecular weight of 4 - 6x106 g/mol and a 61 dosage level of 50 ppm; ii) changing pH had a minor effect; and iii) a slightly, lower moisture content was observed at a pH close to a neutral. Y u and Attia (1991) first tested the use of a totally hydrophobic latex in coal dewatering. They used a FR-7A latex and a high air pressure filter apparatus for the filtration of a -40 um Pittsburgh No. 8 raw coal. Their results showed that the filtration rate increased with the flocculant concentration, and the cake moisture content decreased under an air pressure of 20 psi for the tested coal fines. 3.3.3 Filtration Enhanced by Surfactants There is a wider divergence of opinions regarding the mechanisms of dewatering enhanced by surfactants than those by the other materials, such as oils or flocculants. The surface tension and viscosity lowering mechanism was suggested by Silverblatt and Dahlstrom (1954), Nicol (1976) and Lewellyn (1982). But Phillips and Thomas (1955), Keller et al. (1979) and Owen (1984) questioned it. The divergence of these opinions was obviously caused by the variety of surfactants and coals used in their studies. The simplistic model often used to demonstrate the mechanism of dewatering improvement by surfactants is to treat a porous bed of particles as an assembly of cylindrical capillaries. For analysis, a single capillary, which is described by equation (14), can be considered. From the capillary theory, Keller (1979) proposed a two-fold effect of surfactants. Firstly, surfactants effectively lower liquid surface tension. In terms of overcoming the capillary forces in a bed of small particles, this interfacial tension is 62 proportional to the work required to displace the liquid. Secondly, similar compounds have been used by the mineral industry as flotation collectors. In this case, the molecules are adsorbed by a mineral and render it more hydrophobic, which also leads to a reduction of the capillary pressure (pc). The actual dewatering pressure (Ap=p-pc ) across the filter cake increases, thus improving the dewatering process. However, whether the second effect will take place or not depends on the types of surfactants and the solid surface properties. If the surfactants are not adsorbed by the solid, they actually only reduce the liquid surface tension. Sodium dodecyl sulfate, tested by Nicol et al. (1976), can adsorb onto coal and increase its surface hydrophobicity. A surfactant molecule consists of a polar group and a nonpolar hydrocarbon radical. The polar group is hydrophilic in character while the nonpolar group is hydrophobic. Therefore, the orientation of adsorbed molecule at the coal/water interface is another important issue. Gillmore and Wright (1952) were among the first to consider the role of surfactants in coal dewatering. The surfactant Tergital N P G (nonylphenyl polyoxyethylene ether) was used in their studies. They found that the addition of this nonionic surfactant did not change the drainage rates during the initial stages of their laboratory drainage tests when most of the drainable water is removed and when the moisture distribution pattern is being established. The final moisture content of the fine coal filter cake could be reduced from 26.0% to 20.8%. They reported an overall moisture content reduction of about 5%. 63 Silverblatt and Dahlstrom's (1954) experimental data for -2.24 mm fine coal showed that when using Aerosol O T (sodium 2-ethylhexyl sulfosuccinate) as a wetting agent, the effect of surface tension on cake moisture content for concentrations of the wetting agent not greater than 0.02 wt% was relatively small. However, when relatively high concentrations of the surfactants Aerosol O T and Tergitol C W were used, the cake moisture dropped sharply. From their data, they deduced an empirical equation which showed that viscosity was considerably more effective than surface tension in reducing the moisture content of a fine coal filter cake when using relatively low concentrations of the wetting agent. Their suspicion that surface tension lowering played an important role was based on their results showing that moisture contents at 72 and 31 mJ/m 2 surface tension were essentially equal, but a large drop in moisture content occurred between 32 and 26 mJ/m 2 . They concluded that such a small difference in surface tension could not be the cause for all of the moisture reduction but that some kind of surface adsorption involving the surfactant played a major role. Phillips and Thomas (1955) agreed with Silverblatt and Dahlstrom. Conclusions drawn from their own experiments were that the increased rates in the flow of water through porous beds observed upon heating might be due to surface-tension decrease, but they primarily attributed the improvement to the more marked decrease in the viscosity. Based on their experiment in which an unidentified wetting agent was used to reduce surface tension by half, but at the same time the contact angle of coal surface was reduced from 45° to 2 5 ° and final moisture content was not affected, they pointed out that the surface-tension reduction effect was not helpful in reducing the cake moisture. Sometimes they observed that it had a detrimental effect on cake moisture reduction. 64 The surfactant Teepol was used in Grey's (1958) dewatering study. His encouraging results showed that when the surfactant was used on a relatively coarse fraction of Betteshanger coal (-30 +60 mesh), the final cake moisture, at high suction, was reduced from 8 to 4 %. But results on finer coal samples were not as good. The reduction of cake moisture was very limited, possibly because of a tendency of the wetting agent to deflocculate the coal and clay and to clog the large pores. According to Grey, a reduction of liquid surface tension is an obvious way of improving dewatering. However, he did not provide any data supporting such a conclusion. Dolina and Kaminskii (1971) employed a variety of surfactants which included cationic, anionic and nonionic surfactants and different coals to study the effect of surfactant types, coal rank and degree of oxidation on coal dewatering. They found that there was no clear correlation between the cake moisture content and surface tension lowering, and that whether the moisture content increased or decreased, depended on the type of surfactant used. Their results also indicated that moisture content was a function of coal rank: coals at both ends of the scale retained more moisture in the cake than those of intermediate rank. Oxidation was detrimental to the dewatering process. They concluded that the used anionic surfactants were not effective, however, the tested cationic and nonionic materials were among the best of the identified dewatering additives for both oxidized and metallurgical coals. Later, Nicol (1976) selected dodecylammonium bromide and sodium dodecyl sulfate for dewatering -500 +76 um coarse fraction of fine coal. From his experimental data, he suggested that the more pronounced effect of the anionic surfactant on coal 65 dewatering compared to the cationic surfactant could be explained in terms of adsorption at the solid solution interface. The anionic surfactant reduced cake moisture from 12% to 5% at dosages ranging from 0 - 1 0 0 ppm, while the cationic surfactant reduced moisture from about 12% to 11.7% at the same dosages. Although no direct measurements of the surfactant adsorption on coal have been reported, electrophoretic data suggested that the coal surface was negatively charged at neutral pH. Consequently, the concentration loss from solution arising from adsorption was likely greater in the case of the cationic surfactant, and therefore, higher concentrations were required to produce a given equilibrium surface tension than in the case of the anionic surfactant. These conclusions are in contrast to Dolina and Kaminskii's observation. Keller et al. (1979) carried out the most fundamental studies on the application of surfactants in fine coal (-500 um) dewatering. From their experiments, it is obvious that surfactants play a two-fold role by adsorbing at both the liquid/air and solid/liquid interfaces. The former leads to a decrease in surface tension and affects filtration in accordance with the capillary theory. The latter results in a more hydrophobic solid. They proposed the models for surfactant adsorption on coal surfaces which are shown in Figure 3.3.3-2. 66 ADSORBED ORIENTATION OUTER HELMHOLTZ PLANE (+) © ® - > - — ^ INNER HELMHOLTZ PLANE (-) Q H - J j J ^ ^ ^ V O 11" H n H 9 9 9 9 COAL MINERAL COAL COAL (+) (-) (+) jl jl G 9 Q 9 9 9 e e e . i j , J e e . v C U A 1 — | MINERJ\L~COA"L—I COAL CH2 ( + ) (") (+) C H 2 Figure 3.3.3-2 Models for the adsorption o f anionic and cationic surfactants on coal According to these models, a reduction in moisture content is due to the hydrophobic character imparted to the solid surface by surfactant adsorption. According to Keller et al.'s study, adsorption of an anionic surfactant on coal surface takes place in the Stern layer in such a fashion that the hydrophobic hydrocarbon group extends towards the aqueous phase. If this is true, one would assume that a cationic surfactant molecule would adsorb on these surfaces with its hydrophilic head towards the aqueous phase due to the electrostatic repulsion between the head and the same charged coal surface and hence would be much less effective. But their results show that the cationic surfactants also give good reduction in cake moisture. To explain the behavior of cationic surfactants, they assumed that adsorption of cationic surfactants occurs in the outer Helmholtz plane over a layer of hydroxyl ions, which are adsorbed by the coal surface. This will make the coal surface more hydrophobic and cause the observed decrease in the cake moisture content. However, no great differences in effectiveness were found between the best anionic and the other tested surfactants on the particular coal tested. Finally, Keller et al. (1979) pointed out that good dewatering aids should be characterized by the following features: 67 1) The surfactant molecules should adsorb on the coal and orient with the hydrocarbon tails extended to the water phase. This makes the coal surface more hydrophobic. 2) The hydrocarbon tail should be as hydrophobic as possible, and still retain a degree of water solubility of the whole molecule. 3) The head group should be the least hydrophilic group possible, and still retain the solubility of the whole molecule. 4) The adsorbed molecule should have a large cross-sectional area so that fewer molecules are required to affect the interfaces. 5) The surfactant should decrease interfacial tension at the liquid-air interface as much as possible at low concentrations. 6) The molecules should not be prone to hemi-micellation or, i f they are, it should occur at a sufficiently high concentration, which can be avoided. Keller et al.'s results were later confirmed by Gala (1982), who also tested a variety of surfactants in the dewatering of -500 urn fine coal samples. Fine coal dewatering can be improved by using surfactant solutions and heating. This technique was examined by Baker and Deurbrouck (1976). A n approximately 9% moisture content reduction for -35 mesh coal slurry over that of an untreated filter cake was observed. However, it is difficult to assess whether this reduction in cake moisture content was due to viscosity reduction, surface-tension lowering, or other effects. Brooks and Bethell (1979) also studied the use of surfactants in coal filtration. Their results indicated that a cationic surfactant (aliphatic diamine) could improve the dewatering of the filter cake. A minimum cake moisture content was obtained at about 0.4 kg/t, with a 10% reduction in moisture content observed. However, the cake moisture content increased remarkably at amine dosages exceeding 0.5 kg/t. The probable reason for this was the formation of cracks in the filter cake. Like Keller's study, the hemimicelle concept was also used to elucidate the surface phenomena that could improve the dewatering at lower dosages and impair the dewatering at higher dosages. New commercial surfactants A C C O A L 4335 and A C C O A L - D R I 1000 were tested by Lewellyn (1988) for the dewatering of-0.59 mm, -0.59+0.15 mm and -25 +0.59 68 mm coal fractions. The efficient moisture content reduction from about 15% to 11% for a -0.59 mm fraction at a dosage of 400 g/t was demonstrated and the foaming problems caused by other surfactants in the plant circuit were solved. Ofori et al. (1989) used a large number of proprietary surfactants, including four anionic, one cationic and ten nonionic surfactants, to study their efficiency in enhancing the filtration of a number of Australian coals. In terms of the effect of coal size distribution, their results showed that relatively fine coal (with the highest proportion of fines) did not respond favorably to all of the surfactant types. With respect to the effect of surface tension, dewatering enhancement did not correlate well with the lowering of the liquid surface tension. However, the rank of the tested coals appeared to have a significant effect on dewatering when surfactants were used. The dewatering of the lower rank coals responded well to all the tested reagents; the higher rank coal dewatering was improved more by cationic than by other surfactants. The cationic surfactants performed well on all the tested coals. This result contrasts with that obtained by Nicol (1976), but is in agreement with Dolina and Kaminskii (1971). Groppo and Parekh (1994) recently studied the surface chemical control of fine coal to reduce moisture content. The presence of 50 ppm copper (Cu + 2 ) or 10 ppm aluminum (Al + 3 ) ions lowered cake moisture from 32% to 28%. The cake moisture reduction could be attributed to the coagulation of fine particles along with precipitating hydroxides from the solution. Addition of 0.2 kg/t anionic surfactant along with either 50 ppm copper or 10 ppm aluminum ions reduced cake moisture to 22.8%. A dosage of 0.1 kg/t cationic surfactant provided a 23%> moisture filter cake. 3.3.4 Filtration Enhanced by Hydrophobization As analyzed in the preceding section, coal hydrophobicity can affect the dewatering process; it influences particle aggregation, which in turn affects the capillary structure and permeability of the filter cake, resulting in a reduction of capillary pressure. 69 It is well established that the more hydrophobic the coal, the easier water is displaced by air from the cake. This point has been well recognized by many researchers in this area. In Gray's (1958) early fundamental studies, E S S O medium fuel oil and liquid paraffins were used as oily additives in the dewatering of -600 pm coal. About 5% - 7% in moisture content reductions were reached at an oil dosage of 5 - 20 kg/t. Gray also added oil to the slurry in the presence of surfactants and found dewatering to be even more effective. In fact, this oil was even more effective than some water soluble flocculants. He explained that this might be due to the fact that, in addition to the agglomeration effect, oils also increased the receding contact angle of water on the coal surface. By changing the hydrophobicity of glass spheres with chlorosilane, Phillips and Thomas (1955) found the water retained in the pendular state to be 11% for a 0° contact angle and 1.2 % for a 90° contact angle. But in Harris and Smith's (1957) study, no decrease in the final moisture content was reported. Only the pendular state in a bed of glass spheres was reached more rapidly by increasing the contact angle of the glass spheres. They observed that the pendular state occurred when a sufficient amount of water had been removed from the bed so that no continuous network of water existed except discrete rings of water at the points of contact of the particles. Nicol and Rayner (1980) studied the oil-assisted filtration of fine coal. The oil used was commercially available household kerosene, and the coal was 96.5% passing 0.5 mm of Australia B H P fine clean coal. Their results showed a considerable cake moisture content reduction from 18 % to 11% with an oil addition of 1% to the tested coal. Evans and co-workers (1984) applied a hydraulic piston type filter to dewater a -100 mesh fine coal slurry with the addition of pitch fiber (shredded waster newsprint). Their experimental results indicated that the cake moisture was reduced from 9.4% without dose to about 7% at dosages of 5% pitch and 3% fiber. Loo and Slechta (1987) performed oil-assisted -500 pm coal filtration experiments in a pilot plant. About 1.5% diesel oil addition caused a cake throughput increase by about 40%. Their investigation showed, however, that the cake moisture 70 content also slightly increased when the throughput was raised with oil addition. Increases of both the cake moisture content and the filter cake throughput with the addition of flocculant Sanfloc AH200P were also observed in the pilot plant. Dewatering improvement by oil agglomeration has been studied by Capes et al. (1989), who did extensive work in this field. The results of their detailed study may be summarized as follows: an important aspect of oil agglomeration for moisture reduction is its inherent dewatering capability due to the displacement of water by the attached oil; and oilconcentration is found to be the most.significant parameter in coal dewatering by agglomeration. They generalized the complicated relationships between agglomerate size, moisture content and oil concentration. The concentration regions, designated as low, intermediate and high, correlate with the physical characteristics of the agglomerates. A low oil concentration, (less than 5%), in the pendular region, produces a flocculated coal product with a high moisture content. The consolidated floe structure traps much of the suspending liquid. Intermediate concentrations of oil (5%-30%), in the funicular and capillary regions, yields microagglomerates and large discrete agglomerates, and the moisture content of the recovered coal begins to decrease more significantly. The reason for such behavior in this region is that agglomerates grow rapidly in size and the total surface area of the agglomerates decreases correspondingly, as does the surface moisture. Moreover, the internal moisture is reduced because the oil fills the agglomerate voids and displaces the water. While higher oil concentrations result in a coal-oil amalgam, the moisture content of the products begins to increase. Figure 3.3.4-1 illustrates the effect of oil concentration on agglomerate size and moisture content in the oil agglomeration of fine coal. Capes et al. (1989) also proposed an equation to correlate product moisture with agglomerate size. 71 discrete / | coal-oil agglomerates / j amalgam o c a) c Z3 Oil concentration Figure 3.3.4-1 The effect of oil concentration on agglomerate size and moisture content (Capes 1982) In their study of the separation of pyrite from coal by oil agglomeration, Good et al. (1994) showed that because of coal heterogeneity, water is always present between the agglomerated solid particles. The effect of dewatering aids on coal wettability is another important aspect which was studied by some researchers (Owen et al. 1984 and Schwendeman et al. 1972). Owen's studies showed that coal surface contact angle decreased with the addition of commercial polyacrylamide flocculants to various Eastern U.S. coals. Schwendeman (1972) also provided direct contact angle evidence for such behavior in the presence of 15 surfactants and water soluble polymers. In summary, the application of hydrophilic flocculants (polyelectrolytes) in the filtration and dewatering of -500 urn fine coal has been studied quite extensively. Although surface tension may play a role in moisture reduction, increasing the solid/liquid contact angle seems to be the most important factor in the filter cake moisture content reduction by surfactants. In addition, with respect to the effect of coal particle size, flocculants reduce the cake moisture of coarse particles more effectively than fine particles. 72 Much less research has been conducted on the use of oils in coal dewatering perhaps because oil addition tests revealed the ineffectiveness in promoting coal dewatering: the high cost resulted from the high oil consumption. As is apparent from the literature review, research on the use of the three types of additives (flocculants, surfactants and oils) in ultrafine coal dewatering has not yet produced conclusive results. Even the effect of coal particle size distribution on its dewatering in the presence of additives is not well understood. 73 CHAPTER 4 MATERIALS 4.1 Coal Samples The three coal samples were obtained from Fording River mine and Line Creek mine in British Columbia. Their wettability differs widely. The Ford-4 metallurgical coal sample is very hydrophobic, Ford-13 is quite hydrophilic and the Line Creek-7 (LC-7) sample lies somewhere in-between these two (Palmes & Laskowski 1993). Their wettability and other characteristics are described in Table 4.1-1 (Laskowski et al. 1991). Table 4.1-1 Assessment of coal surface wettability and other related properties Ford-4 L C - 7 Ford-13 Demineral ash (%) 0.14 0.12 0.55 Total acidity (Meq/g) 0.04 0.16 1.22 Contact angle * (deg) 113 81 20 Equil. moisture (%) 0.94 1.73 4.59 Advancing contact angle measured not on demineralized but on as received coal. The measurements were conducted on pellets made under pressure from fine coal. 4.1.1 Coal Sample Preparation for Filtration Tests The three coal samples received from the mines were first crushed below 13 mm and cleaned wet on a concentrating table. The concentrates were dried at 50 °C and then crushed to minus 2 mm. The coal samples were finally pulverized and ground in a rod / 74 mill for three to four hours to below 45 um. Proximate analyses of the three coal samples are given in Table 4.1.1-1. The size distributions for the Ford-4, Ford-13 and Line Creek-7 ultrafine coal samples were measured with an ELZONE - 280 particle size analyzer and are shown in Table Apdx - 1 in Appendix. From a calculation shown in Apdx-16, the specific surface area for the Ford-4 ultrafine coal is 1.54 m2/g. Table 4.1.1-1. Proximate Analyses of Ford-4. Ford-13 and Line Creek-7 Coal Samples Sample Ash (%) Total Moisture (%) Volatile Mat. (%) *p (2 dramf (%) Ford-4 8.37 0.71 19.74 78.3 Ford-13 9.42 2.84 22.43 74.4 L C - 7 8.95 1.18 20.28 77.7 F C d m m f - Fixed carbon content on the dry and mineral matter free basis The specific surface areas of the -45 um ultrafine coal and the -500 pm fine coal samples were measured using a QUANTA-SORB apparatus and the BET method. For nitrogen at 196 K, three adsorption measurements in the range of 0.1 - 0.3 relative pressure P/Po (where P is the equilibrium pressure; and Po, the saturation pressure of the adsorbate at the adsorption temperature) were taken to calculate the surface area using the BET equation. Values of 1.72 m2/g for the Ford-4 ultrafine coal and of 270 cmVg for the Ford-4 fine coal were obtained. The surface electrical properties of the coal particles were characterized by electrokinetic measurements. An electrophoresis apparatus (II - U V A ) was employed. A solution of 10 2 M KCI was utilized as a supporting electrolyte. The zeta potentials measured at various pHs for the Ford-4, Ford-13 and LC-7 coal samples are shown in Figure 4.1.1-1. 75 40 Figure 4.1.1-1 Zeta potential vs. pH curves for the tested coal samples In order to investigate the differences in filtration behavior between -45 urn ultrafine coal and -500 um fine coal with the addition of various chemical additives, the -500 pm Ford-4 and Ford-13 coal samples were prepared from the clean coal by crushing and grinding and used to represent respectively hydrophobic and hydrophilic fine coal samples. The size distributions of these two fine coal samples are given in Tables Apdx - 2 and Apdx - 3 (Appendix). 4.1.2 Coal Sample Preparation for Hydrophobic Agglomeration of Ultrafine Coal The as received Ford-4 run-of-mine coal was crushed and then wet-ground to below -45 um in a rod mill for 3 hours. This sample was used in the study of hydrophobic agglomeration of run-of-mine ultrafine coal with the use of hydrophobic latices, 76 hydrocarbon oils and oil emulsions. The size distribution of the sample was analyzed using an ELZONE-280 particle size analyzer. The ash content of the sample was assayed at 17.29%. Its particle size distribution is shown in Table Apdx - 4 (Appendix). The same ultrafine clean coal sample which was used in beneficiation of the silica/coal mixture and the clay/coal mixture was employed in the filtration tests. 4.1.3. Sample Preparation for Abstraction of Latex and Oil Droplets In order to reduce the effect of residual ultrafine particles in the supernatant on transmittance of the solution, minus 600 jam coal particles and silica particles were deslimed by screening out ultrafine particles that are smaller than 45 urn. The -600+45 urn fraction of Ford-4 coal and the silica particles are used in tests of abstraction kinetics and isotherms of latex and oil droplets by coal and silica. The particle size distributions for the coal and silica samples are shown in Tables Apdx - 5 and Apdx - 6 (Appendix). 4.2 Silica Sample Minusil Silica was used to carry out experiments on the separation of ultrafine coal and silica by hydrophobic agglomeration processes with the use of hydrophobic latices, hydrocarbons and oil emulsions. Its particle size was measured using the E L Z O N E - 280 particle size analyzer; the mean size (d50) of the particles was found to be about 2 um. 4.3 Clay Sample Kaolin was utilized in the experiments on the separation of an artificial mixture of ultrafine coal and clay particles by hydrophobic agglomeration processes, which included 77 beneficiation with the use of hydrophobic latices and oil emulsions. The top size of the clay particles was found to be 10 pm and the mean size (d50) was about 3 pm. The ash content of the sample was measured to be 85.67%. 4.4 Chemical Reagents In general, the solutions used were prepared by weighing out a predetermined amount of the chemical and diluting it with distilled water. When aging was a problem, fresh solutions were always prepared as required. Distilled water was used throughout the study. 4.4.1 Polyelectrolytes Polyacrylamide (PAM), a commonly used flocculant in the coal industry with a molecular weight of 6 x 106 g/mol, was obtained from Polysciences, Inc. It is a hydrophilic polyelectrolyte. 4.4.2 Hydrophobic Latices The UBC-1 and FR-7A totally hydrophobic latices were used in the hydrophobic agglomeration processes and filtration/dewatering processes. The FR-7A latex was provided by Calgon Corporation, Pittsburgh. The UBC-1 latex was synthesized in the Department of Mining and Mineral Process Engineering at The University of British Columbia. The FR-7A latex was examined under a Hitachi S-500 Scanning Electron Microscope and the size of the latex particles was found to range from 0.06 to 0.1 pm. The film prepared by evaporating FR-7A latex turned out to be very hydrophobic with a 78 60° advancing contact angle. Wettability of the film prepared with the U B C - 1 latex was similar to that prepared with the FR-7A. The size of the UBC-1 latex particles was found to range from 0.5 to 1.0 pun. As electrophoretic mobility measurements indicate (Figure Apdx-2, Appendix) (Laskowski et al. 1994), these latex particles are negatively charged over the whole p H range due to the presence of anionic groups. As a result, these latex particles form a very stable aqueous colloidal system. The particles of the U B C - 1 latex are very different from the hard spheres of polystyrene emulsified with potassium oleate, and are instead very soft, flexible and easily deformed upon drying. Like the F R - 7 A , the U B C - 1 latex also belongs to the group of totally hydrophobic agglomerants. 4.4.3 Semi-hydrophobic Flocculant Polyethylene oxide (PEO), a non-ionic polymer, supplied by Aldrich Chemical Company, Inc. with a molecular weight of 4x106 g/mol, was selected as a semi-hydrophobic flocculant. 4..4.4 Surfactants 4.4.4.1 Cationic Surfactants A number of surfactants were selected to emulsify oily hydrocarbons or were directly used as filtration additives. Dodecyl amine hydrochloride ( D D A HC1), dodecyl amine (DDA) and C , 6 - C 1 8 long chain amines were selected as cationic surfactants. Dodecyl amine hydrochloride was supplied by Eastman Kodak Company, dodecyl amine was obtained from Aldrich Chemical Company Inc., and C 1 6 - C 1 8 long chain amines, trade name Armeen H T D , were kindly provided by A K Z O Chemicals Ltd. 79 Dodecyl amine hydrochloride was dissolved in water and a 0.1% by weight stock solution was prepared. Dodecyl amine (1% wt) was used to dissolve in kerosene or other oils. A predetermined amount of kerosene (or other oils) containing D D A was then emulsified in 0.1% wt hydrochloric acid solution. C 1 6 - C , 8 long chain amine was either used as an aqueous solution or dissolved in kerosene. When it was utilized in an aqueous solution, a C 1 6 - C i 8 long chain amine stock solution (0.1 % wt) was prepared according to the information provided by A K Z O . 0.5 gram of C 1 6 - C , 8 amine was dispersed in 50 ml of distilled water with agitation at 70 °C. Glacial acetic acid (0.105 ml) was then introduced and the agitation was continued until the solution became homogeneous. The solution was finally diluted 10 times to be used as a stock solution. The stock solution was heated and shaken prior to being utilized. 4.4.4.2 Anionic Surfactants Sodium dodecyl sulfate and sodium cetyl sulfate, provided by Polysciences Inc. and Research Plus Inc., were selected as anionic surfactants in this study. They were prepared as 0.1%) wt aqueous solutions prior to being used. 4.4.4.3 Non-ionic Surfactants Ethoxylated nonylphenols (CO -520 , CO-610 and CO -730) , supplied by G A F Corporation, New York, USA, were utilized as non-ionic surfactants. These aqueous-soluble surfactants were prepared as 0.1% wt aqueous stock solutions before being utilized. 80 4.4.5 Oily Hydrocarbons Kerosene and several ESSO heavy oils such as ESSO-806, ESSO-846, ESSO-1156, ESSO-2300, ESSO-2600 and ESSO-904 were obtained from ESSO Petroleum Canada. They were used as oil agglomerants and filtration additives in the form of emulsions, or were used as received. ESSO-806 and ESSO-846 have the lowest densities among the tested ESSO heavy oils. ESSO-2600 is a relatively high density oil with a high viscosity and therefore, it is very difficult to disperse in an aqueous solution. ESSO-904 is a mixture of ESSO EX-466 and ESS0-1156. ESSO Ex-466, a catalytic cracking bottom product, is a black, asphalt-like oil. ESSO-1156 oil is a light fuel oil with low viscosity. By mixing the two and measuring the density of the mixture, it was found that ESSO-904 oil consists of 55% ESSO Ex-466 and 45% ESSO-1156 by weight. Figure Apdx-17 in the Appendix is a typical flowsheet of one refinery operation showing the steps by which the ESSO oils are produced. Finished Base Oils include lubricating oils, ESSO-2600 and 2300, greases, etc. Table Apdx-14 in the Appendix lists typical physical properties and chemical compositions of ESSO oils. 4.4.6 Dispersing Agents Sodium hexmetaphosphate (SHMP), Na 6 P 6 0 1 8 , supplied by BHD Chemicals Ltd., England, was used as a dispersant in the hydrophobic agglomeration of ultrafine coal with the use of hydrophobic latices and oil emulsions. Sodium tripolyphosphate (STP) was also utilized as a dispersing agent in this study. The chemical structures of these dispersants are shown below: 81 f f f 0 0 o Sodium tripolyphosphoate V ° N X X y > 0 + 6 Na" / P N Sodium hexmetaphosphoate °x° o o 4. 4. 7 Regulators Dilute hydrochloric acid and sodium hydroxide solutions were used as p H regulators. 82 CHAPTER 5. EXPERIMENTAL TECHNIQUES 5.1 Adsorption and Attachment Measurements 5.1.1 Abstraction of a hydrophobic latex by coal and silica particles The coal and silica samples used in the abstraction tests with the UBC-1 hydrophobic latex and oil droplets are described in Chapter 4. For each test, 20 grams of the coal or silica were used. The abstraction test was carried out in a beaker equipped with four baffles on the inside wall. A standard six-paddle stirrer with speed control and paddle height adjustment manufactured by Phillips & Bird Company was employed. The 20 gram sample was mixed with distilled water to make a 900 ml suspension. A predetermined amount of the hydrophobic latex was added to the slurry and stirring was maintained for a certain length of time to allow the latex to attach to the solid particles. After 5 minutes of sedimentation, the aggregates settled down to the bottom of the beaker, and the upper part of the solution could be decanted and centrifuged (8,000 rpm for 3 minutes). Subsequently, the concentration of latex in the supernatant was determined as described in section 5.1.2. 83 5.1.2. Determination of Hydrophobic Latex Concentration All materials are capable of scattering light to some extent due to the Tyndall effect. The noticeable turbidity associated with many colloidal dispersions is a consequence of intense light scattering (Shaw 1980). The turbidity of a colloidal system is proportional to the concentration of dispersed particles in solutions. The turbidity of a colloidal system is then defined by the expression: I t/I 0 = exp[-xx] (17) where I0 is the intensity of the incident light beam, I, is the intensity of the transmitted light beam, % is the wave length of the sample and x is the turbidity. For a specific sample, % is a constant, and I0 can be fixed during measurement for a given spectrophotometer. As a result, the relationship between x and Ln (It) becomes linear. Therefore, the concentration of hydrophobic latex in solutions can be determined based on the turbidity, or transmittance (T), of the solution. In these measurements, the transmittance is used instead of turbidity ( T = 100 - x ). The solution transmittance was measured at a wavelength of 800 nm against distilled water with a Perkin-Elmer Lambda UV - VIS spectrophotometer. The latex concentrations were determined using a calibration curve. The differences between the original and residual latex concentrations were taken as the abstracted amount of latex. 84 100 UBC-1 Latex Concentration (ppm) Figure 5.1.2-1 Calibration curve for UBC-1 latex concentration against solution transmittance. The solution transmittance was measured at a light wavelength of 800 nm The calibration curve shown for the UBC-1 latex was obtained by using known concentrations of the latex in solutions. The latex concentrations in the abstraction tests were determined from the measured transmittance and the calibration curve (Figure 5.1.2-1). The relationship between the transmittance T (%) and UBC-1 latex concentration C (ppm) can be expressed as: Ln (T) = 4.584- 0.041 C (18) For equation (18), the standard coefficient error is 0.00081, the residual squared is 0.996, and the standard estimated error of Ln (T) is 0.0292. 85 A s far as reproduciblity of the results is concerned, the abstraction data for the hydrophobic latex by coal are very consistent. The spectrophotometric measurements of transmittance of latex solutions are given in Table Apdx-10. The average points for three readings are plotted graphically in Fig. 5.1.2-1. As Table Apdx-10 demontrates the spread of the results for each concentration is extremely small as shown by a very small standard deviation (< 0.02). The r 2 values (>0.99) close to 1 show that the calibration relationship between the transmittance and the UBC-1 concentration is fairly close to linear (insert in Fig. 5.1.2-1). 5.1.3 Adsorption of Surfactants The adsorption of a surfactant from an aqueous solution onto coal particles can be obtained by measuring the concentrations of the surfactant in the aqueous solution before and after the adsorption. From the change in surfactant concentration, the amount of surfactant adsorbed by the coal particles can be determined from the following equation: N = (c, - c2) / m where N = the number of moles of surfactant adsorbed per unit weight of adsorbent under the same conditions as in filtration (mole/g); c,, c 2 = the surfactant concentrations in the coal /water suspension and in the filtrate (mole/1); m = the weight of adsorbent (coal) (g). In the adsorption tests, the initial surfactant concentration in the coal/water slurry was known, and the surfactant ( D D A or DDS) concentration in the filtrate was determined using an ion surfactant selective electrode (Rippin & Laskowski 1985, Castro 86 and Laskowski, 1985). The selective electrode was made by gluing a piece of PVC membrane to the bottom of a hard PVC tube with a THF solution of PVC as an adhesive. The membrane responsive to dodecyl ammonium ions (DDA +) was prepared in accordance with the procedure used by Maeda et al. (1981) and Hayakawa et al. (1982) from a mixture of 0.35 g PVC, 1.00 g of bis (2-ethylhexyl) phthalate plasticizer and 10 mg of an active complex. The active complex was prepared by dissolving equivalent amounts of dodecyl amine hydrochloride (DDA.HC1) and sodium dodecyl sulfate (SDDS) in distilled water. The resultant solutions, having pHs between 4 and 5, were heated until almost boiling, and were then combined. A colloidal suspension and crystal growth was noticed as the resulting solution cooled. The white precipitate of DDA+DDS" was washed with distilled water, followed by a double recrystallization from acetone. The calibration curves showed that the responses of the electrode (mV) to D D A HC1 and SDDS (M) were nearly linear, especially in the D D A + and DDS - concentration range from 10"5 to 10"2 M . The test results also indicated that when surfactant concentrations were high (i. e., 5xl0"4 to 10"2 M), the effect of pH on the potential of the ion-surfactant selective electrode could be ignored. However, when the surfactant concentration was low (i. e., less than 5x10-4 M), the effect of pH on the potential had to be considered. Therefore, in this adsorption study the solution pH was controlled in the range of pH 7 to pH 8 throughout the measurements. The experimental set-up is sketched in Figure 5.1.3-1. The actual surfactant concentration in the filtrate can be found from the calibration curve. 87 electrode electrode PVC tube reference solution special membrane tested solution Fig 5.1.3 -1. Sketch of an experimental system for the measurement of surfactant concentration with the use of ion-surfactant selective electrodes Figures 5.1.3-2 and 5.1.3-3 show the calibration curves for dodecyl amine chloride (DDA) and sodium dodecyl sulfate (DDS). The adsorption results for surfactants on coal turned out to be very sensitive to p H . The first two adsorption trials failed due to the p H fluctuations during the tests. The p H must be well controlled at a certain consistent level. Relatively speaking, the reproducibility of the results obtained at low surfactant concentrations is lower than those obtained at high surfactant concentrations. Based on the test results, when p H fluctuations are controlled within ± 0 . 0 1 , standard deviations for the repeated tests are quite small. As Table Apdx-9 shows, standard deviations of electrode potential readings at high concentrations (e.g. 100 mg/1 ) of dodecyl amine chloride are smaller (about 0.15 mV) than those at low concentrations (e.g. 0.01 mg/1) of dodecyl amine chloride (about 1.44 mV). Reproducibility of average electrode potentials at a high surfactant concentration is also much better than that at a low concentration (see Table Apdx-9). 88 001 0.1 1 10 100 Dodecyl amine concentration (mg/l) Figure 5.1.3-2 Calibration curve showing electrode potential versus concentration of DDA.HC1 0.01 0.1 1 10 100 Sodium dodecyl sulfate concentration (mg/l) Figure 5.1.3-3 Calibration curve showing electrode potential vs. concentration of DDS 89 5.1.4 Abstraction of Oil Droplets by Coal Particles The method which was used to measure the abstraction of the latex was also employed to investigate the abstraction of oil droplets by coal particles. The -500 + 45 pm clean coal sample and the same standard six-paddle stirrer were used. Before the test, a calibration curve was established by using known concentrations of oil-droplets in solutions. After abstraction and decantation, the spectrophotometer was used to determine the transmittance of the supernatant. The oil droplet concentration then could be determined from the measured transmittance by using a calibration curve. The amount of oil droplets abstracted by the coal particles was calculated based on the following equation: D = (C, - C 2 ) / m where D - the amount of oil attached per unit weight of coal (mg/g); C, , C 2 - the oil droplet concentrations in the solutions prior to and after the abstraction (mg/1); m - the weight of the adsorbent (coal) (g). The calibration curves for the kero-DDA and kero-DDS emulsified oils versus transmittance are shown in Figures 5.1.4-1 and 5.1.4-2. 90 As in the case of the hydrophobic latex, the results for the abstraction of oil-droplets by coal are very consistent. Table Apdx-11 presents the results used to plot graphically the calibration curves for concentration of the kero-DDA emulsion versus transmittance (Fig. 5.1.4-1). The standard deviation for the transmittance measurements is bleow 0.05. The r2 values (>0.99) close to 1 show that the calibration relationship between the transmittance and the kero-DDA concentration is fairly close to linear. Similar reproducibilities were obtained for kero-DDS and these are included in the Thesis. 100 (D O c ro I w i= ro kerosene emulsified in water with addition of 1 wt% of dodecyl amin The solution was measured at a light wavelength of 800 nm Transmittance Log Scale 100 80 60 40 Ln (T) = 4.595 - 0.041 C 1^  = 0.9996 0 200 400 600 800 1000 Kero-DDA Oil Emulsion Concentration (ppm) 100 200 300 400 500 600 700 800 900 1000 1100 Kero-DDA Oil Emulsion Concentration (ppm) Figure 5.1.4-1 Calibration curve of kero-DDA oil emulsion concentration versus transmittance 91 100 1200 Oil Concentration (%) Figure 5.1.4-2 Calibration curve of kero-DDS oil emulsion concentration versus transmittance 5.2 Emulsification of Oily Hydrocarbons To obtain a stable oil-in-water emulsion, surface active substances have to be used. To obtain a stable emulsion, simple shaking, high shearing and PIT (phase inversion temperature) emulsification techniques were tested. The emulsification of oil by simple shaking was performed in a bottle (300 ml) containing 0.01% emulsifier, 1% oil and 99% distilled water. The bottle was shaken for one minute on a shaker. In the emulsification of oil by high shearing, a commercial high speed blender with speed control was employed. For each test, 250 ml of emulsion was produced, which 92 contained 0.001-0.01 % emulsif ier by weight o f emuls ion , 1% o f o i l and 9 9 % o f d i s t i l l ed water. E m u l s i f i c a t i o n was carried out at 30,000 r p m for 3 minutes. The effect o f the emuls i f ica t ion temperature o n the mean diameter o f o i l droplets i n O / W type emuls ions is w e l l documented (Shinoda and Saito 1969). W h e n a non- ion ic surfactant is used as an emulsif ier , a f inely dispersed emuls ion can be obtained i f the system is i n i t i a l l y emuls i f ied close to the phase invers ion temperature (PIT) (about 2 - 4 °C b e l o w the P I T appears opt imal) o f the emulsif ier . A t this temperature the interfacial tension is sma l l ( M a n d a n i and Fr iberg 1978) and the o i l droplets are f ine ly dispersed i n the solut ion. The emuls i f ica t ion process is subsequently fo l l owed by rap id c o o l i n g to r o o m temperature, at w h i c h the coalescence rate is s low. The c o o l i n g process adds s m a l l droplets f rom the incipient phase separation o f the surfactant phase (Fr iberg and Solans 1978). S u c h an emuls i f ica t ion process is referred to as emuls i f ica t ion by the P I T method. The P I T temperatures for ethoxylated nonylphenols C O - 5 2 0 and C O - 7 3 0 were found to be 60 °C and 61 ° C , respectively. A c c o r d i n g l y , emulsif icat ions w i t h these surfactants and kerosene were carr ied out at 57 °C and 58 ° C , respectively. The o i l emuls ions produced i n a h i g h speed blender w i t h the addi t ion o f surfactants (amines, sulfates or ethoxylated nonylphenol) are very stable. The size d is t r ibut ion o f emul s ion droplets (e.g. k e r o - D D A and the emul s ion obtained by emuls i fy ing kerosene i n a cetyl sulfate solution), measured 24 hours after emuls i f ica t ion , is the same as that measured from a freshly prepared emuls ion . The size d is t r ibut ion measurements o f kerosene droplets emuls i f ied i n a 2 .7x10-4 M cety l sulfate so lu t ion (Table A p d x - 1 2 i n the A p p e n d i x ) indicate that the size and size dis t r ibut ion o f kerosene 93 droplets measured after 24 hrs are almost the same as those measured from the emulsion freshly produced. 5.3 Assessment of Coal Surface Wettability Coal surface wettability was evaluated by means of contact angle measurement. The receding contact angles were determined using a Rame-Hart contact angle goniometer with a controlled environment chamber. Lumps of coal were selected, cut into 1" x 2" x 1/2" size, encapsulated in an epoxy resin to prevent fragmentation during handling, and cut on one side to expose the coal surface using a saw. The coal surface was then polished flat using an abrasive disk down to 600 grit and stored under distilled water until using. The prepared lump was conditioned in the flocculant solution with each predetermined concentration for 5 minutes under standardized hydrodynamic conditions, and then transferred into the chamber together with the flocculant solution. The measured surface was positioned upside down, and a bubble was introduced onto the surface using a special syringe. Each reported contact angle value was the average of the contact angle values taken from six point measurements. 5.4 Measurement of Particle Size Distributions The particle size distributions of the -500 urn fine coal were obtained by screening. The E L Z O N E 280 particle size analyzer was used to measure the size 94 distributions of the -45 um ultrafine coal and the size distributions of the oil droplets produced by emulsification. For the size analysis of the -45 um coal particles, a few grams of the sample were placed in a beaker with four baffles on the wall and mixed with some water by stirring for 30 minutes in order to well disperse the coal particles in water. A few drops of the slurry were syringed and then diluted in a beaker containing 300 ml of a 6% NaCl solution. A few drops of the diluted slurry were used as the sample for the size analysis. The emulsions were prepared as described in the preceding section. For each oil droplet size analysis, a small amount of the emulsion was taken and diluted in a 6% NaCl solution. 5.5 Electrokinetic Tests Zeta-potentials of the coal particles and the oil-droplets were measured using a II-U V A electrophoretic apparatus. For each test, a certain amount of coal fines was mixed in 10 - 2 M KC1 electrolyte solution. The emulsified oil was diluted and zeta potentials of the oil droplets were also measured in a 10 -2 M KC1 solution. The coal particles were placed in an electrophoresis cell with two electrodes. A voltage was applied between the two electrodes to produce an electric field in the cell and the charged particles responded by moving toward one of the electordes. The electrophoretic mobility of the particles is proportional to the magnitude of the zeta-potential. The sign of the charged particles was determined from the direction of their movement. The zeta-potential of the particles is 95 available from the provided calculation table or can be calculated by the Smoluchowski equation (Shaw 1980) based on the measured electrophoretic mobility of the particles. The electrokinetic measurements with coal particles and oil droplets are very reproducible. Each zeta-potential value is determined as an average of at least 6 readings. Standard deviation for each set of six zeta-potential measurements is less than 2.5 m V . 5.6 Flocculation and Hydrophobic Agglomeration A commercial high speed blender (20,000 rpm) was used to disperse the coal particles in distilled water with or without the addition of sodium hexmetaphosphate (SHMP). Flocculation and hydrophobic agglomeration for the filtration experiments were carried out in a standard six-paddle stirrer with speed control and paddle height adjustment from the Phillips & Bird Company. A beaker with four baffles on the wall was used as a slurry mixing container. For each test, 18 grams of coal were mixed with water to make a 900 ml suspension. The suspension dispersed in a high speed blender was introduced to the beaker, and stirring was maintained at 350 rpm for at least 20 minutes. At the end of this period, a very diluted flocculant (or a hydrophobic latex dispersion) was slowly added into the beaker. Stirring was continued for a further 2 minutes at a predetermined high speed and then reduced to 50 rpm for 1 minute, to promote floe growth, then stopped. After a 3 minute flocculation process, the upper 500 ml of suspension was siphoned out. The flocculation percentage was calculated using the following formula: (%) F = {[(M/18)100-44]/56}100 96 where M is the mass of dry solids (in grams) in the bottom 400 ml of the suspension after a 3 minute settling period. The detailed flocculation process for the filtration tests is shown in Figure Apdx - 3 (Appendix). Initially, flocculation and agglomeration tests were repeated twice or more. The results produced under the same conditions were found to be quite comparable. Not many discrepancies appeared during these tests. Therefore, most of the later flocculation and agglomeration tests were conducted only once. In the beneficiation experiments using a hydrophobic latex, the cleaning of the ultrafine run-of-mine coal or the separation of the mixture of clean coal and silica (or clay) was performed using the same set-up as used to carry out flocculation for the filtration tests. In these tests, a dispersant had to be used. After selective hydrophobic agglomeration, the agglomerated slurry was transferred onto a 400 mesh screen to separate the coal agglomerates from the dispersed gangue. 5.7 Oi l Agglomeration Oil agglomeration also belongs to the family of the hydrophobic agglomeration processes. It is described separately since the test procedures are somewhat different from those of hydrophobic flocculation or agglomeration. The coal sample used in these tests was minus 45 urn ultrafine run-of-mine coal that was prepared by grinding in a rod mill. The 50/50 mixtures of minus 45 um ultrafine clean coal and silica, and the 50/50 mixtures of coal and clay were also tested. For each test, 400 ml of a 5 wt.% coal water slurry sample was subjected to the standardized test 97 procedure using a commercial high speed blender with speed control. The pH of the slurry was first adjusted to a desired level. In some tests, SHMP or STP dispersant solution was added to the slurry and conditioned for 2 minutes. A predetermined amount of oil or emulsified oil was then injected into the slurry and followed by high speed mixing at 20,000 rpm for 30 seconds. The mixing speed was then reduced to 1,000 rpm for another 30 seconds to allow the agglomerates to grow. At the end of agglomeration, the slurry in the blender was transferred onto a 400 mesh screen. The agglomerates were retained on the screen as a clean coal product while the mineral particles dispersed in the aqueous solution passed through the screen. The clean coal and refuse products were dried, weighed and assayed. 5.8 Filtration and Dewatering A funnel type filter, made of a plexiglass cylinder 50 mm in diameter and 150 mm in height with a scale on the wall, was used as a vacuum filtration and dewatering device. A vacuum pump was employed as the vacuum source. A moisture trap, a vacuum gauge and a bleeder were used to adjust the filtration vacuum. The flocculated or agglomerated slurry, transferred from the beaker into the filter, was subjected to filtration and dewatering under suction. When water disappeared from the cake surface, the vacuuming was continued for three minutes and then the cake was taken out of the filter unit under suction. It was weighed prior to and following drying in an oven at 110 °C to determine the moisture content of the cake. The filtration rate was measured by recording the time 98 taken for the liquid level to pass between two calibrated marks on the filter wall. The filtration and dewatering system is sketched in Figure Apdx - 4 (Appendix). Table Apdx 13 provides information on reprodubility of the filtration results. The average of relative experimental deviation (Rdev) is ±4.5% for the filtration rate and ±2.7% for the cake moisture content with the use of flocculants ( P A M and PEO), latices (UBC-1 and FR-7A), or emulsified oils. When surfactants are utilized in the filtration tests, the relative experimental deviations are slightly higher, +6.2% for the filtration rate and ±4,7% for the moisture content. Relative experimental deviation Rdev is defined as: Rdev = (x; - X) / X , where x ; is a single test result and X is the average value of a group of repeated test results. 5.9 Surface Area Determination The surface area of the coal particles was measured by means of the Q U A N T A -SORB instrument, which uses the principle of gas (N2) adsorption. One gram of coal sample was used for each measurement. 99 5.10 Definition of Measurements in Beneficiation Tests To evaluate test results, some measurement terms, such as flocculation percentage, product ash content, ash reduction, product combustible recovery, and Gaudin index, are often used in beneficiation tests. The flocculation percentage is already defined in section 5.6, and the rest of terms are defined as below: 1) Ash content is determined based on the standard of ash content analysis by burning 1 gram sample of coal in a muffle furnace to temperatures of 700 to 750 °C, and calculated using the following formula: Ash content (%) = (wt. of ash residual after burn / wt. of original sample) x 100 2) Ash reduction is determined by the following equation: Ash reduction (%) = [(A f - A c ) / A f ]x 100 where A f is feed coal ash content, A c is clean coal ash content. 3) Combustible recovery is defined by the following equation: Combustible recovery (%) = Yield (%) x [(100 - Ac)/(100 - A f)] where yield is the yield of coal product, which is calculated by: Yield (%) = (wt. of coal product / wt. of coal feed) x 100 4) Gaudin index is used to evaluate the selectivity of some reagent or process. The larger the Gaudin index value, the higher the selectivity of the reagent or process. It is defined as Gaudin index = [(100 - A c ) x A, ] / [(100- A t ) x A J where A, is ash content of tailings, A c is the same as defined before. 100 C H A P T E R 6. R E S U L T S A N D D I S C U S S I O N 6.0 Introduct ion This chapter is structured as follows. Sections 6.1 and 6.2 primarily focus on selective flocculation and hydrophobic agglomeration with the use of flocculants and hydrophobic latices that include water soluble flocculants (polyelectrolytes), semi-hydrophobic flocculants and totally hydrophobic latices. Section 6.3 describes the results of oil agglomeration of ultrafine coal, including i) comparison of oil agglomeration using emulsified oils with conventional oil agglomeration; ii) effect of various parameters on the agglomeration of ultrafine coal; and iii) the role of surfactants in oil agglomeration. Section 6.4 focuses on the filtration/dewatering of clean coal products from the beneficiation processes. Therefore, this section mainly deals with the effects of the tested flocculants, surfactants, and emulsified oil, which were used in the beneficiation processes, on the filtration and dewatering of ultrafine and fine coal. 101 6.1 Total and Selective Flocculation of Ultrafine Coal 6.1.1 Abstraction of a Hydrophobic Latex by Coal Particles The standard calibration curves for the UBC-1 latex are shown in Figure 5.1.2-1. The coal and silica samples used in the abstraction tests are described in Chapter 4. Figure 6.1.1-1 shows the kinetics of abstraction of the UBC-1 latex by the coal and silica particles. The initial concentration of UBC-1 latex in these kinetic abstraction tests is 30 mg/1. The abstraction rate of the latex by the coal is initially rapid and reaches a plateau after about 30 minutes. As these results indicate, about 30 minutes is necessary to reach an equilibrium. The results also reveal that the amount of the latex abstracted by coal particles is much higher than that abstracted by silica at the same pH. The abstraction rate of the latex by the coal is more than 15 times higher than that by the silica. Consequently, it can be concluded that the hydrophobic latex has a much higher affinity towards the hydrophobic coal surface than silica. It can, therefore, be expected that when the hydrophobic latex is added to a mixture of coal and silica particles, it will preferentially attach to the coal particles. 102 * * . * — — Coal Silica 3 e 0 o Time (min) Figure 6.1.1-1 Abstraction kinetics of UBC-1 latex by Ford-4 coal and silica particles at pH 6.4 (Initial UBC-1 latex concentration 30 mg/1) 20 UBC-1 Latex Concentration (mg/1) Figure 6.1.1-2 Abstraction of UBC-1 latex by Ford-4 coal (1996 new sample) and silica at pH 6.4 103 Abstraction of the U B C - 1 latex by the coal and silica are shown in Figures 6.1.1-2 and 6.1.1-3. The abstraction time under stirring is set for 15 minutes. A l l the abstraction curves show the same general form with each curve approaching saturation abstraction at about the same concentration ( « 30 mg/1). As Figure 6.1.1-3 shows, the amount of abstracted latex in each case initially increases with increasing equilibrium concentration of the latex and then remains constant. Only a negligible amount of latex was found to attach to the silica particles. 25 Coal at pH4.2 o Coal at pH6.4 Coal at pHl 0.3 pH4.1 pH6.5 H I 0.51 20 30 UBC-1 Latex Concentration (mg/1) 50 Figure 6.1.1-3 The effect of pH and latex concentration on the abstraction of UBC-1 latex by Ford-4 coal (1996 new sample) and silica particles In addition, these abstraction test results also indicate a very strong effect of p H on the abstraction. As shown in Figure 4.1.1-1, the i.e.p. of Ford-4 coal is 7.3. The electrophoretic measurements (Figure Apdx-2, Appendix) indicate that the U B C - 1 latex particles are negatively charged over the whole p H range. At pHs below the i.e.p. of the coal, the coal surface is positively charged. In addition to the hydrophobic interaction, 104 apparently, electrostatic attraction forces also favor the abstraction of the negatively charged latex particles onto coal at pHs < i.e.p. At pHs > i.e.p. the coal surface assumes a negative charge, and although abstraction still takes place, it diminishes because of the electrostatic repulsion between the coal surface and the latex. For hydrophilic silica, the abstraction is so low that the effect of pH cannot be detected. 6.1.2 Effects of Flocculant Type and Coal Wettability on Flocculation and Hydrophobic Agglomeration In this series of tests, three ultrafine coal samples, Ford-4, LC-7 and Ford-13, were used. Their surface wettability increased in the following order: the Ford-4 is the most hydrophobic, the Ford-13 the most hydrophilic, and the LC-7 lies somewhere between these two. The effects of UBC-1 dosage on the flocculation of the three tested coals are shown in Figure 6.1.2-1. As in the case of the abstraction of UBC-1 by coal, with an increase in the UBC-1 dosage the flocculation of the Ford-4 coal initially improves significantly. At a dosage of 100 g/t, flocculation reaches a plateau. The flocculation of the LC-7 coal reaches a plateau around a dosage of 200 g/t, whereas UBC-1 is essentially unable to flocculate the oxidized Ford-13 coal. These results confirm that the UBC-1 is a totally hydrophobic latex which by its nature is the most effective in flocculating hydrophobic particles (e.g. Ford-4 coal). The flocculation effectiveness of UBC-1 is lower for less hydrophobic particles. When coal particles are highly wettable by water (Ford-13), the UBC-1 latex does not tend to attach to such particles, which means that the abstraction of the UBC-1 latex by the hydrophilic coal is very low, probably similar to that by silica. Figure 6.1.2-2 shows the effect of polyethylene oxide (PEO) on the flocculation of the three different coals. Similar to the UBC-1 latex, PEO does not flocculate the Ford-13 105 coal. P E O was previously tested by Gochin et al. (1985) and it was shown to effectively flocculate hydrophobic coal only. These results confirm Gochin et al.'s conclusion. However, P E O effectively flocculates the Ford-4 hydrophobic coal at low dosages, and the L C - 7 at higher dosages. The floes formed by P E O are much smaller and weaker than those formed by the U B C - 1 latex. z o o o o FORD-4 200 FORD-13 300 400 DOSAGE (g/t) 500 700 Figure 6.1.2-1 Effect of UBC-1 latex dosage and coal wettability on hydrophobic flocculation of ultrafine coal at pH 6.8 100 FORD-4 FORD-13 100 200 300 400 500 600 700 PEO DOSAGE (g/t) Figure 6.1.2-2 Effect of PEO dosage and coal wettability on flocculation of ultrafine coal at pH 6.8 700 PAM DOSAGE (g/t) Figure 6.1.2-3 Effect of PAM dosage and coal wettability on flocculation of ultrafine coal at pH 6.8 100 0 100 200 300 400 500 600 700 DOSAGE (g/t) Figure 6.1.2-4 Comparison of UBC-1 with FR-7A latex at pH 6.8 107 The effect of polyacrylamide (PAM) dosage on the flocculation of the three ultrafine coals is shown in Figure 6.1.2-3. These results clearly indicate that P A M does not exhibit any selectivity. A l l the coals, irrespective of their surface properties, are flocculated. This is not surprising since P A M was developed for flocculation in general solid/liquid separations and should flocculate all suspended particles. Therefore, it is impossible to use P A M in fine coal beneficiation. In the case of P A M , the curves show a clear optimum which was not observed when the UBC-1 hydrophobic latex or PEO semihydrophobic flocculant was used. To compare UBC-1 with FR-7A latex, the same three coals were utilized in the flocculation tests. The results are presented in Figure 6.1.2-4, and reveal that at the same dosages, the flocculation percentages are generally higher for UBC-1 than for FR-7A. The floes appeared to be larger and stronger with the use of UBC-1 than with FR-7A. 108 6.1.3 Effect of pH on Flocculation and Hydrophobic Agglomeration The effect of p H on flocculation was tested at a constant P A M concentration of 200 g/t and a constant U B C - 1 latex concentration of 200 g/t. The flocculation of Ford-13 oxidized coal with U B C - 1 was found to be negligible over the whole p H range (Figure 6.1.3-1). P A M flocculates both Ford-4 and Ford-13 coals, and the effect of p H is similar for the both cases. It is known that P A M contains carboxylic and amide groups. The electrostatic attraction between the positively charged coal and the negatively charged carboxylic group definitely facilitates the adsorption of P A M onto Ford-4 coal at pHs below 7.3, and on Ford-13 at pHs below 4. At neutral p H (6.4 - 7.0), P A M still flocculates the Ford-4 coal very well as this coal is still positively charged or close to its i.e.p. P A M also flocculated oxidized Ford-13, probably through hydrogen bonds between the amide groups and oxygen atoms on the coal surface, in spite of the electrostatic repulsion between the coal surface and the P A M macromolecules. However, at high pHs, although the hydrogen bondings between the coal and P A M still result in the flocculation of Ford-4 coal particles, and to some extent Ford-13 particles, the repulsion forces between the negatively-charged coal surface and the negatively-charged macromolecules prevail and reduce the flocculation of both coals. Figure 6.1.3-2 Effect of hydrodynamic condition on flocculation (UBC-1 dosage: 200 g/t, P A M dosage: 200 g/t, Ford-4 ultrafine coal) 110 6.1.4 Effect of Hydrodynamic Conditions The bridging flocculation of coal particles with a polymer is generally considered to take place in three stages: i) dispersion of the polymer in the solution, ii) adsorption of the polymer at the interface, and iii) the collision of particles partially covered with the polymer to form bridges. It is obviously impossible to complete the first and third stages without some sort of stirring. In other words, some mixing must be provided for flocculation to take place. The effect of stirring rate on the hydrophobic flocculation of Ford-4 ultrafine coal with the use of the UBC-1 latex and on flocculation with P A M is shown in Figure 6.1.3-2. No flocculation was observed without prior conditioning with either the UBC-1 latex or P A M . With an increase in stirring rate, the flocculation of coal particles with both latex and P A M initially improves dramatically, and reaches maximum at around 100 rpm. At this point, the formed floes appear to be large and weak. With further increase of the stirring rate, the hydrophobic agglomerates with UBC-1 become smaller but stronger. Both types of the floes are able to withstand relatively severe hydrodynamic conditions. The maximum flocculation range is maintained until the stirring rate reaches 300 rpm. Too intense stirring, however, is detrimental to flocculation and hydrophobic agglomeration. 6.1.5 Summary and Discussion The hydrophobic latex, developed in the last decade, is one of the most promising new reagents for the beneficiation of ultrafine coal. The UBC-1 latex is a very good hydrophobic agglomerant for the beneficiation of ultrafine coal. Due to the presence of a I l l surfactant with anionic groups, the latex particles are negatively charged and form a stable aqueous colloidal system. The latex is, however, very hydrophobic (the contact angle on a film prepared from UBC-1 is 60 degrees). The abstraction isotherms for the UBC-1 hydrophobic latex show that it has a much higher affinity towards coal than to hydrophilic silica particles. The abstraction of the hydrophobic latex is very sensitive to solution pH. In an acidic environment, the abstraction densities of U B C - 1 on coal are relatively high. Under alkaline conditions, the abstraction of the latex by the coal is substantially reduced due to the electrostatic repulsion between the coal and latex, and accordingly, the hydrophobic agglomeration of the coal particles was depressed, but selectivity improves. Figure 6.1.5-1 shows the effect of pH, and of the U B C - 1 latex, on the hydrophobic agglomeration of the Ford-4 and Ford-13 coals, as well as zeta potential vs. p H curves for these two coals and the UBC-1 latex. As Figure 6.1.1-3 reveals, the U B C - 1 latex particles easily attach to the surface of the hydrophobic Ford-4 coal over a broad p H range. The abstraction of the latex starts decreasing only around p H 10. The abstraction curves correlate extremely well with the hydrophobic agglomeration results (Figure 6.1.5-1); Ford-4 coal is agglomerated by the U B C - 1 latex over a broad p H range from 3 to 9 in which the zeta potential of the U B C - 1 latex particles is negative. For the Ford-4 coal, the zeta potential is positive in an acidic environment and negative in the alkaline range (the values change from +30 m V down to about -20 mV). In this whole range the Ford-4 coal particles are agglomerated by the U B C - 1 latex. Since the i.e.p. for the Ford-13 coal is around p H 4, the zeta potential for this coal assumes more negative values in alkaline solution, about -40 m V at p H 9. However, even around p H 4-5 when the zeta potential 112 values for the Ford-13 coal particles are close to zero, the UBC-1 latex does not agglomerate this hydrophilic coal. This indicates that the classical D L V O theory, which takes into account only electrical and van der Waals forces, cannot explain these observations. The Ford-4 coal suspension is agglomerated even around pH 8-9 when both the latex and the coal particles are negatively charged, and this must be due to hydrophobic forces. On the other hand, the Ford-13 coal is not agglomerated by the U B C -1 or FR-7A latex even in the pH range 3-5 where zeta potential values for the latex and the coal are close to zero. Thus, the good agglomeration of the hydrophobic Ford-4 coal, and the negligible agglomeration in the tests with the hydrophilic Ford-13 coal can only be explained by taking into account these structural forces. This situation resembles closely the system studied by X u and Yoon (1989). The UBC-1 latex was also tested for flocculating other inherently hydrophobic minerals such as graphite and molybdenite. The tests revealed that the latex could flocculate graphite at high dosages but was unable to flocculate molybdenite. However, the latex was found to flocculate molybdenite well in the presence of MIBC (Castro and Laskowski 1997). This indicates that the attachment of the latex to solid surfaces is likely due not only to hydrophobic interactions and electrostatic forces, but probably also involves some specific interactions imparted by the polar groups on the latex and solid particles. Figure 6.1.5-1 Zeta potential of coal particles and UBC-1 latex particles, and hydrophobic agglomeration of Ford-4 and Ford-13 coals using UBC-1 latex 114 From a comparison of the reported flocculation studies with P E O , it appears that this polymer is a good flocculant for those solids which are hydrophobic to some extent. The fact that P E O poorly flocculated the Ford-13 hydrophilic coal particles is in line with the findings of Gochin et al. (1985), who concluded that P E O flocculated only hydrophobic anthracite but not oxidized anthracite. Simpson (1990) and Lekili (1990) measured P E O adsorption onto coal and graphite and showed that the amount of P E O adsorbed onto graphite was higher than on anthracite since the graphite was more hydrophobic than anthracite. However, their experiments on the adsorption of P E O on anthracite and oxidized anthracite indicated that there was still some P E O adsorbed on the oxidized anthracite although the amount of adsorbed P E O on the oxidized coal was about 70% less than that on the unoxidized coal. Rubio and Kitchener (1976) found that P E O could adsorb on methylated silica surface but not on hydrophilic silica. A n anionic water-soluble flocculant, P A M in water looks like a long stretched macromolecule with numerous branches that contain its amide and carboxyl groups (Akers and Ward, 1995). The interactions of P A M with the coal surface may involve hydrogen bonding, dispersion forces and/or electrostatic forces. At low and neutral pHs, as shown in Figure 6.1.3-1, good flocculation of coal particles with P A M was obtained. In the acidic and neutral p H range, the Ford-13 coal is negatively charged, but the zeta potential is low. Therefore, P A M can still flocculate this coal because of the interactions via hydrogen bonding. However, in an alkaline environment, the carboxyl groups are fully ionized, and the flocculation of Ford-4 coal was substantially reduced. P A M did not flocculate Ford-13 particles in alkaline media, probably due to extremely high Coulombic repulsions. 115 6.2 Beneficiation of Ultrafine Coal using Hydrophobic Latices 6.2.1 Predetermination of Slurry Solid Content for Selective Flocculation Preliminary tests revealed that the solids content of slurry had a significant effect on the ash content of the concentrate and the coal combustible recovery. At a very low solids content, combustible recovery was low, though ash rejection and selectivity were high. In contrast, an overly high solids content resulted in a poor separation selectivity but a high combustible recovery. When solid content is lower than 2%, the combustible recovery falls off. Therefore, 2% solid content was used in the selective flocculation tests (Figure 6.2.1-1). Figure 6.2.1-1 Effect of slurry solids content of Ford-4 ultrafine coal on hydrophobic agglomeration 116 6.2.2 Effect ofDispersant on Selective Flocculation Several dispersants, including sodium hexametaphosphate (SHMP), sodium silicate, and sodium tripolyphosphate were tested at dosages ranging from 0 to 1200 mg/l. S H M P provided the most satisfactory dispersion of the solid particles. The results are shown in Figure 6.2.2-1. A predetermined amount of ultrafine coal was added to the beaker and water was then filled to 900 ml to form a 2% solid content slurry. A specific amount of dispersant was added to the slurry, and followed by stirring for 10 minutes. After 3 minutes of sedimentation, the solid content in the beaker was determined to find the best dosage, at which the maximum dispersion percentage was achieved. The dispersion percentage at a given dispersant dosage is defined as: Dispersion (%) = (solid weight in the upper 50% volume)/(Original solid weight in this part) 100, . 20 -o * • 0 200 400 600 800 1000 120 SHMP D O S A G E (mg/L) Figure 6.2.2-1 Dispersion tests with sodium hexametaphosphate using Ford-4 ultrafine coal 117 0 50 100 150 200 250 300 350 Dosage of SHMP (mg/l) Figure 6.2.2-2 Effect of dispersant (SHMP) dosage on hydrophobic agglomeration of a mixture of 90% ultrafine Ford-4 coal and 10% kaolin with UBC-1 latex (UBC-1 dosage: 400 g/t, natural pH of 6.4, shear rate: 320 rpm, Ford-4 sample) The results indicate that the optimum dosage of S H M P is 200 to 300 mg/1, at which a maximum dispersion percentage of 84% is obtained (Figure 6.2.2-1). The effect of S H M P dosage on hydrophobic agglomeration of a mixture of 90% coal and 10% kaolin with the U B C - 1 latex is presented in Figure 6.2.2-2. From the trends of the three curves, it is evident that the maximum ash reduction and the best selectivity (Gaudin Index) are reached at a S H M P dosage of 300 mg/1. S H M P does not seem to have any effect on coal combustible recovery in the tests. 118 6.2.3 Separation of Coal from Coal-Silica Mixture by Hydrophobic Agglomeration Effects of latex dosage, p H and stirring rate on the beneficiation of ultrafine coal from a 50/50 mixture of coal and silica at 2% solid content are shown in Figures 6.2.3-1 to 6.2.3-3. Two hydrophobic latices (UBC-1 and FR-7A) and P A M were utilized in these experiments. As shown in Figure 6.2.3-1, the ultrafine coal could be selectively agglomerated from the mixture by UBC-1 latex. Ash content was substantially reduced from 53.9% (feed) to 14.9% (product) in a one-stage separation at over 95% combustible recovery. At low dosages, relatively small and weak agglomerates were observed along with poor selectivity and a low combustible recovery. When latex dosage increased to 300 - 400 g/t, the high combustible recovery and high ash reduction resulted in a very high separation selectivity, indicated by the Gaudin index which reaches a maximum. The results also showed that the performance of the F R - 7 A latex was similar to the U B C - 1 with slightly lower selectivity and coal recovery. In contrast to the hydrophobic latices, P A M showed an extremely poor separation selectivity: the water soluble P A M flocculated almost all the suspended solid particles regardless of their surface properties. Figure 6.2.3-2 shows the effect of p H on beneficiation with the two hydrophobic latices. Generally speaking, for both latices, low pHs produced relatively large and strong agglomerates with high ash contents, while alkaline pHs turned out to produce relatively small and weak floes but with low ash contents. The best selectivity occurred at around 119 neutral p H o f 7-8, at w h i c h h i g h combust ible recoveries and l o w ash agglomerates were obtained. A s shown i n F igure 6.2.3-3, the best separation select ivi ty was achieved at a s t i r r ing rate o f about 300 rpm. It was found that large and loose floes were usual ly generated at l o w st i rr ing rates, whereas sma l l and strong floes were produced under re la t ively intense m i x i n g condi t ions . The b i g and loose floes not on ly entrapped a large amount o f ash but also contributed to a l o w combust ible recovery since such weak floes c o u l d not wi ths tand manipu la t ion dur ing the separation process. Howeve r , when the st irr ing rate became too h igh , the intense turbulence destroyed the pre-formed large floes and resulted i n part ial agglomerat ion and sma l l agglomerates. Hence , a c lean product w i t h l o w ash and l o w combust ib le recovery was produced. A l l the results demonstrated that the separation select ivi ty w i t h U B C - 1 was better than that w i t h F R - 7 A . 120 Figure 6.2.3-1 Selective hydrophobic agglomeration tests of a 50:50% coal/silica mixture as a function of flocculant dosage 40 h-z LU 35 1-Z 30 o HC 25 CO < 20 o FLO 15 DOSAGE: 350 G/T 300 RPM F-4:SILICA=1:1 T o " T 1 DOSAGE: 350 G/T 300 RPM F-4:SILICA=1:1 JBC^  7 — S 5 TT5 Ti FEED ASH CONTENT: 53.94 % UBC-1 FI DOSAGE: 350 G/T 300 RPM F-4:SILICA=1:1 -e-T T — T PH T 1 Figure 6.2.3-2 Selective hydrophobic agglomeration tests of a 50:50% coal/silica mixture of as a function of pH 122 Figure 6.2.3-3 Selective hydrophobic agglomeration tests of a 50:50% coal/silica mixture of as a function of stirring rate (rpm) 123 6.2.4 Separation of Coal and Kaolin by Hydrophobic Agglomeration According to some reports (Anon., 1972), coal mineral matter primarily consists of silica and kaolin minerals, which generally make up over 70% of the total ash content. Therefore, the separation of a coal-clay mixture was also investigated. Two samples of ultrafine coal and kaolin with different ratios, 90%: 10% and 50% : 50%, were used in these tests. Only the U B C - 1 latex was utilized since the previous experiments had already demonstrated that U B C - 1 was the superior agglomerant. The selective agglomeration of the coal-kaolin mixture with the latex was also tested as a function of latex dosage, slurry p H and hydrodynamic conditions. The results are shown in Figures 6.2.4-1 to 6.2.4-3. After a one-stage separation, a concentrate with 11.8% ash was obtained at 98% recovery from the 90:10 mixture of coal and clay. The initial ash content of the mixture was 17.6%. The optimum latex dosages were in the range of 500 to 700 g/t, and the best p H range was around p H 6 - 7. Beyond p H 8, the combustible recovery and the Gaudin Index were significantly reduced (Figure 6.2.4-2). From the Gaudin Index plot as a function of stirring rate, the best stirring rate was found to be at 200 rpm. The results for the beneficiation of the 50:50 mixture of coal and clay closely followed the trends obtained with the 90:10 mixture, but lower combustible recoveries (slightly over 80%>) throughout the whole tested dosage range were obtained. However, the ash reduction was surprisingly high for such a high clay content; the ash was reduced from 47.3%) in feed to 15.4% in the product. 124 50 (Natural pH~6.2, stirring rate: 320 rpm) 50-50% of kaolin and coal 90-10% of coakkaolin 100 200 300 400 5flfl 5fl0 700 8flfl 9oo _100 & 90 > o o CU Cd 05 E o o 80 S 70 60 50 - e r =5= 50"-50% of kaolin andcoaT 90-10% of coakkadrin 100 200 300 400 500 600 700 800 9&0 100 x cu "2 80 60 > o $ 40 ? 20 co CD 90-10% of coal:kaolin 100 200 300 400 500 600 700 800 900 Dosage (g/t) Figure 6.2.4-1. The effect of UBC-1 dosage on hydrophobic agglomeration of coal-kaolin mixtures (Natural pH6.2-6.4, stirring rate 320 rpm, SHMP dosage 200 mg/l) 125 PH Figure 6.2.4-2. The effect of pH on hydrophobic agglomeration of coal-kaolin mixtures (UBC-1 dosage 300 g/t, stirring rate 320 rpm, SHMP dosage 200 mg/l) 126 22, 20 18 o O 16 ID < 14 12 10| 100, £-90 (U §801 I 70| w i i 60| E o " 5 0 v50-50% of kaolin:Coal -e- 9U-1U% ot coal:kaotin "roxr 200 •w 600 9M0%OTcoal:kaoltri 700 50-50% of kaolin:Coal ^ T 0 0 2 0 0 300 400 500 60Tj TOO Stirring Rate (rpm) Figure 6.24-3. The effect of stirring rate on hydrophobic agglomeration of coal-kaolin mixtures (UBC-1 dosage 300 g/t, natural pH6.2-6.4, SHMP dosage 200 mg/1) 127 6.2.5 Beneficiation of Ultrafine Run-of-Mine Coal by Hydrophobic Agglomeration In these tests using the latex, the ultrafine Ford - 4 run-of-mine coal with an ash content of 17.3% was utilized. The beneficiation results with U B C - 1 latex are presented in Figures 6.2.5-1 to 6.2.5-3. In this one-stage beneficiation process under the optimum conditions, i. e., p H of 7-8, stirring rate around 300 rpm, and latex dosage 300 - 400 g/t, a clean product with 8.1%> ash content could be produced on a screen at over 95%> combustible recovery. Figure 6.2.5-2 shows the strong effect of p H on the beneficiation results. In acidic solutions, the high ash content of the product indicates that the hydrophobic latex almost loses its selectivity due to strong electrostatic interactions between the latex, coal, and mineral particles as well as probably high entrainment. At p H 7-8, the selectivity is certainly enhanced since the product ash content is significantly reduced from 15.0% to 8.1%. The selectivity index improves from 7 to 40. However, beyond p H 9, because of the strong electrostatic repulsion, a weak abstraction of the latex by coal was observed. Consequently, a low recovery resulted. The tests on the effect of latex dosage on hydrophobic agglomeration (Figure 6.2.5-1) indicate that the optimum dosages for upgrading the Ford-4 ultrafine coal range from 200 to 400 g/t, at which the ash content of the product almost reaches the minimum level and coal recovery was over 97%. Likewise the previous selective separation tests with the mixtures, 300 rpm was again found to be sufficient for this type of process (Figure 6.2.5-3). In all the beneficiation tests, the agglomerates were separated on a 400 mesh screen and were strong enough to withstand mechanical manipulation during the screening process. 128 0 100 200 300 400 500 600 700 800 900 UBC-1 Dosage (g/t) Figure 6.2.5-1 The effect of UBC-1 dosage on the cleaning of finely ground Ford-4 coal by hydrophobic agglomeration (Natural pH 6.7, stirring rate 320 rpm, SHMP 300 mg/l) Figure 6.2.5-2 The effect of pH on the cleaning of finely ground Ford-4 coal by hydrophobic agglomeration (UBC-1 dosage 400 g/t, stirring rate 320 rpm, SHMP 300 mg/l) 129 Figure 6.2.5-3 The effect of stirring rate on the cleaning of finely ground Ford-4 coal by hydrophobic agglomeration (Natural pH 6.7, UBC-1 dosage 400 g/t, SHMP 300 mg/1) 6.2.6 Summary and Discussion The latex particles produced by emulsion polymerization have 'hairs' protruding from the surface. They are prevented from collapsing back onto the particle due to the presence of charged groups and are about 5 - 1 0 pm long (van de Ven et al. 1983). As these results indicate, a hydrophobic interaction between the hydrophobic coal and hydrophobic latex is obvious. The results of the tests show that selective hydrophobic agglomeration is a very promising beneficiation method for fine coal. A l l the results demonstrate that the new hydrophobic latex agglomerates selectively only hydrophobic particles of bituminous coal. The agglomerates that are produced are strong enough to withstand mechanical 130 manipulation and can be separated on a screen. This new agglomeration process is very similar to conventional oil agglomeration, but the reagent consumption in this process is about 100 times lower than the oil consumption required in the conventional oil agglomeration process. As in the latex abstraction tests, the beneficiation results also reveal a sensitivity to the slurry pH values. The optimum selectivity at high recoveries was always observed over a neutral and slightly alkaline pH range. A better hydrophobic agglomeration was obtained at low pHs, but it was less selective, a typical trade-off in which better selectivity in alkaline media was accompanied by lower recoveries. Under all tested conditions, PAM, a water soluble polyelectrolyte, showed very poor selectivity in the beneficiation process. PAM flocculated all the tested coals and mineral particles regardless of their surface properties. 131 6.3 Oil Agglomeration of Ultrafine Coal with Oil Emulsions 6.3.1. Emulsification of Liquid Hydrocarbons Various techniques are available to prepare oil-in-water emulsions. Such emulsions may be produced by mechanical homogenization using a lab blender for example. However, to obtain a reasonably stable emulsion, a third component - an emulsifying agent must be present during the emulsification process. Dodecyl amine and C 1 6 . 1 8 long chain amine cationic surfactants, sodium dodecyl sulfate and cetyl sulfate anionic surfactants, and polyethoxy nonylphenols non-ionic surfactants were selected for the experiments. 6.3.1.1 Oil droplet size distribution The emulsification procedures are described in Chapter 5. The oil was emulsified in the aqueous phase at a natural p H of 6.4. Table Apdx - 7 (Appendix) summarizes the results of the oil droplet size measurements for a large number of emulsions. The effects of both surfactant type and addition amount on oil emulsification were examined. With respect to the effect of surfactant type on emulsification (Table Apdx - 7, Appendix), the sizes of the oil droplets in all the tested emulsions with the addition of surfactants were much smaller than the size of the oil droplets obtained by mechanical homogenization without any surfactant. Among the tested surfactants, sodium cetyl sulfate looked like to be the most effective. It produced smaller oil droplets with a mean size of 1.34 pm. The dodecyl amine and C 1 6 . 1 8 long chain amine were almost equally effective in the reduction of the oil droplet size as sodium dodecyl sulfate. The non-ionic surfactant, nonylphenoxy polyethanol CO-610, was slightly less effective. 132 Generally speaking, with increased surfactant concentration, the size of the oil droplets was reduced. Under the same emulsification conditions, when the concentration of dodecyl amine was increased by 10 times, the top size of oil droplets was reduced from 5.8 to 4.0 pm, and the mean size was reduced from 1.9 to 1.4 pm. For the long chain amine, as the concentration increased, the size reduction of the oil droplets was more evident. Both the top size and mean size of the oil droplets were clearly reduced. The top size dropped from 9.0 to 3.8 pm and the mean size from 2.8 to 1.4 pm. In addition, changes of the emulsification procedure will also affect the oil droplet distribution. As Table Apdx - 7 (Appendix) indicates, dissolving the surfactant in the oil phase is more efficient than dissolving the surfactant in the aqueous phase, particularly at low surfactant concentrations. The emulsification of kerosene at a low concentration of surfactant in oil (0.1 wt%), is a good example. As shown in the table, the top and mean (d50) sizes, obtained when a 0.1 wt% of C ] 6 - C 1 8 amine is dissolved in oil, approximately equals those sizes obtained when the oil is emulsified in a solution with a 10 times higher G ] 6 . 1 8 amine concentration. The size distributions of the oil droplets for all the emulsions are shown in Figures Apdx-5 to Apdx-15 in the Appendix. The size distributions are skew and the mean size deviates from the center of the distribution, approaching the side of small sizes. This means that the proportion of large size droplets to the total number of droplets is very small. 133 6.3.1.2 Electrokinetic potential of oil-droplets in aqueous solutions Electrokinetic measurements were used to study the surface charge characteristics of the oil droplets. The kerosene emulsions were prepared by either emulsifying kerosene directly in a surfactant solution or first dissolving the surfactant in kerosene and then emulsifying in water. The measurements of zeta-potentials of kerosene droplets are shown in Figures 6.3.1.2-1 to 6.3.1.2-6. As seen from Figure 6.3.1.2-1, the zeta potential of the oil droplets obtained from kerosene emulsified with dodecyl amine can be, depending on p H , either positive or negative. As the amine concentration increases, the zeta-potential values become more positive, and the isoelectric point of the kerosene droplets moves toward higher p H from p H 4.7 up to about p H 11, which is the iso-electric point of colloidal amine (Laskowski et al. 1988). A similar trend can be found in Figures 6.3.1.2-2 to 6.3.1.2-4. Comparing Figure 6.3.1.2-1 with Figure 6.3.1.2-2, it can be seen that, at the same surfactant concentration, the longer chain surfactant ( C 1 6 . l g amine) is more powerful in increasing the positive value of the zeta-potential of oil droplets. Figures 6.3.1.2-3 and 6.3.1.2-4 show zeta-potential as a function of p H for oil droplets produced by dissolving the surfactants into the oil phase and then emulsifying in the aqueous phase. As these results reveal, using the same surfactants at the same dosages, the zeta-potentials are higher when the surfactants were dissolved in oil than when they were added to the aqueous phase. The anionic surfactants sodium dodecyl sulfate and sodium cetyl sulfate were also used in the electrokinetic tests. As shown in Figure 6.3.1.2-5, both dodecyl sulfate and cetyl sulfate shift the original zeta-potential curve of kerosene down to negative values throughout the whole p H range since these are strong electrolytes (the zeta potential - p H curve does not depend on pH). With the use of CO-520 and CO-610 non-ionic 134 surfactants, the original zeta-potential curve for kerosene shifts to slightly more positive values over the whole p H range (Figure 6.3.1.2-6). 150 1 ™ 1 1 1 1 1 1 1 1 0 2 4 6 8 10 12 14 pH Figure 6.3.1.2-1 Zeta potentials of kerosene droplets emulsified in distilled water and in 5.4 x 10"5, 5.4 x 1CH and 5.4 x 10"3 M dodecyl amine solutions 135 150 Figure 6.3.1.2-2 Zeta potentials of kerosene droplets emulsified in distilled water and in 3.9 x 10"5 M and 3.9 x 10"4 M C]6_ 1 8 long chain amine solutions 136 Figure 6.3.1.2-4 Zeta potentials of kerosene droplets and kerosene containing 1% and 0.1% C 1 6 . 1 8 amine emulsified in distilled water 137 -150 • . . . . I 2 4 6 8 10 12 PH Figure 6.3.1.2-5 Zeta potentials of kerosene droplets emulsified in distilled water and in 2.7 x10" 4 M sodium Cetyl sulfate and 3.4 x 10"3 M dodecyl sulfate solutions Figure 6.3.1.2-6 Zeta potentials of kerosene droplets emulsified in distilled water and in 3,7 x 10"4 M CO-520 and 1.7 x 10"4 M CO-610 solutions 138 6.3.2 Abstraction of Oil-Droplets by Coal Particles The abstraction of oil droplets by Ford-4 coal is plotted in Figures 6.3.2-3 and 6.3.2-4. At each concentration, sufficient time and stirring were allowed to ensure that equilibrium was reached. The length of time was predetermined by studying the abstraction kinetics. The rates of attachment are shown in Figures 6.3.2-1 and 6.3.2-2. Figure 6.3.2-1 shows the abstraction kinetics of cationic kero-DDA (kerosene emulsified with the addition of dodecyl amine) oil droplets by coal particles. With an increase in abstraction time, the amount of abstracted oil initially increased rapidly. After about 20 minutes, the abstraction reached a maximum, 200 mg of kero-DDA oil droplets attached to the coal surface, and left a clear final solution. The abstraction behavior of the anionic kero-DDS (kerosene emulsified with the addition of sodium dodecyl sulfate) oil droplets was different. Although the trend of the abstraction of kero-DDS oil droplets by coal looks the same as kero-DDA (Figure 6.3.2-2), the amount of the kero-DDS abstracted on the coal was much lower; only about 16% of the total amount of oil in the system was abstracted. In other words, there was still 84%> of kerosene left in the solution when the abstraction reached an equilibrium. The abstraction of cationic kero-DDA droplets on coal is presented in Figure 6.3.2-3. At oil concentrations below 0.5%), the increase in abstraction density was relatively low. At increasing oil concentration, the attachment density on coal was significantly improved. The maximum abstraction was not reached even at the highest tested concentration. 139 Fig 6.3.2-2 Abstraction kinetics of kero-DDS oil droplets by Ford-4 coal Coal size 600x45 um, oil-droplets concentration 200 mg/1, 350 rpm, natural pH (6.3 140 1400 -1200 I-C o a < 1000 800 600 400 200 2 3 Initial Oil Concentration n Fig 6.3.2-3. Abstraction of Kero-DDA Oil-Droplets by Ford-4 Coal Coal size 0.6x0.045 mm, 20 grams, abstraction time 15 minutes 350 rpm, natural pH (6.4), content of DDA in kerosene 1 wt% 40 Fig 6.3.2-4. Abstraction of Kero-DDS Oil-Droplets by Ford-4 Coal Coal size 0.6x0.045 mm, 20 grams, abstraction time 15 minutes 350 rpm, natural pH (6.4), content of DDS in kerosene 1 wt% 141 The completely different abstraction behavior of the kero-DDS droplets is shown in Figure 6.3.2-4. Apparently, two distinct regions of abstraction are observed. In Region I, the anionic oil droplets carrying DDS molecules are attached to the limited available sites on the coal surface probably through hydrophobic and electrical interactions. At a pH of 6.5 this coal is close to its i.e.p. and initially there are no strong electrical repulsive forces between the coal surface and the negatively charged kero-DDS oil droplets. At a certain abstraction value, the electrical repulsive forces between the negatively charged surface of coal coated by kero-DDS droplets and the oil droplets reach a value that prevents further abstraction. As a result, the maximum attachment was rapidly established; at a relatively low kero-DDS concentration of about 0.25%, the plateau values were reached. In region II, the amount of abstracted kero-DDS on coal was almost constant. Compared to the abstraction of the kero-DDA, the amount of the abstracted kero-DDS by the same coal was found to be much lower. 142 6.3.3 Beneficiation of Ultrafine Coal using Emulsified Oil 6.3.3.1 Oil Agglomeration of Ultrafine Coal with Cationic Emulsion In this series of tests, the cationic emulsion of kerosene, which was produced with kerosene emulsified in 5.4 x lO - 4 M dodecyl amine solution, was used. The finely ground Ford-4 coal that was used in the selective hydrophobic agglomeration was again utilized. However, the solid content of the slurry was 5%, higher than that used in the hydrophobic agglomeration with latices. Prior to oil agglomeration, the suspensions of coal were treated with dispersing agents. The results snowed that this step played an important role in reducing the ash content and improving the combustible matter recovery. The amount of dispersant (100 ppm) used in the oil agglomeration was much lower than that used in the hydrophobic agglomeration. When using the emulsified oil, the agglomeration time was significantly reduced at 30 seconds. In this study, the effects of oil dosage, slurry pH and stirring rate on the beneficiation of Ford-4 raw ultrafine coal were investigated. The results are presented in Figures 6.3.3.1-1 to 6.3.3.1-3. Figure 6.3.3.1-1 shows the effect of dosage of emulsified oil on the ash content of the concentrate, coal recovery and separation selectivity. As seen from the figure, the best separation selectivity was reached at about 2% dosage. At this point, the ash content of the clean coal was the lowest (about 6%), and the combustible recovery was close to 100%. The plot of the product ash versus oil dosage demonstrates that, perhaps 0.25% oil addition is quite sufficient. At this oil consumption, the recovery of combustible matter is over 95%> and the ash content is reduced to about 8.6%. Figure 6.3.3.1-1 Agglomeration of ultrafine Ford-4 raw coal with the use of kero-DDA (Test conditions: pH 7.8, 20,000 rpm, SHMP 100 ppm) Figure 6.3.3.1-2 The effect of pH on the agglomeration of ultrafine Ford-4 raw coal using cationic kero-DDA emulsion (Test conditions: oil dosage 2%, 20,000 rpm, SHMP 100 ppm) SHEAR RATE (x1000 rpm) 75 Figure 6.3.3.1-3 The effect of stirring rate on agglomeration of ultrafine Ford-4 raw coal with the use of cationic kero-DDA emulsion (Test conditions: oil dosage 2%, pH 7.8, SHMP 100 ppm) 146 The effect of pH on the agglomeration process is presented in Figure 6.3.3.1-2. As seen, the optimum results were obtained around pH 8. The effect of impeller speed on the beneficiation results is shown in Figure 6.3.3.1-3. By changing the stirring rate it was possible to achieve better ash reduction and higher combustible recovery. When the stirring rate was increased from 300 rpm (flocculation condition) to 20,000 rpm (oil agglomeration condition), the product ash content was reduced from 10% to 6%, and coal recovery improved from 97.5% to nearly 100%. This figure demonstrates how important hydrodynamic conditions are in the oil agglomeration process. The optimum hydrodynamic conditions for oil agglomeration differ very much from those required in hydrophobic agglomeration or selective flocculation. 6.3.3.2 Oil Agglomeration of Ultrafine Coal with Anionic Emulsion The tested anionic oil emulsion was obtained by emulsifying kerosene in a 3.4 x 10"4 M sodium dodecyl sulfate solution. All the other experimental details were the same as in the preceding section. Figures 6.3.3.2-1 to 6.3.3.2-3 show the beneficiation results obtained in a one-stage separation with the use of the anionic emulsion of kerosene. These figures reveal the same trends (optimum dosage, pH and stirring rate) as observed in the beneficiation results with the use of the cationic emulsion. However, beneficiation with the anionic emulsion is much more selective than beneficiation with the cationic emulsion. At a dosage of 0.5%, a one-stage separation produced a product with 4.8% ash at over 99.5% coal recovery. Even when the oil dosage was reduced to only 0.25%, the recovery of combustible matter was still over 99.5% and the ash content of the clean product was at a 5% level. Figure 6.3.3.2-1 Agglomeration of ultrafine Ford-4 raw coal with the use of a kero-DDS anionic emulsion (Test conditions: pH 7.7, 20,000 rpm, SHMP 100 ppm) Figure 6.3.3.2-2 The effect of pH on agglomeration of ultrafine Ford-4 raw coal using a kero-DDS anionic emulsion (Test conditions: oil dosage 2%, 20,000 rpm, SHMP 100 ppm) 5 10 15 S H E A R RATE (x1000 rpm) 20 25 Figure 6.3.3.2-3 The effect of stirring rate on agglomeration of ultrafine Ford-4 raw coal with the addition of a kero-DDS anionic emulsion (Test conditions: oil dosage 2%, pH 7.7, SHMP 100 ppm) 150 6.3.3.3 Oil Agglomeration of Ultrafine Coal using Non-Ionic Emulsion Non-ionic emulsions, obtained by emulsifying kerosene in a 1.7 x 10"4 M CO-610 and 3.7 x 10"4 M CO520 solutions, were also tested to agglomerate the same ultrafine coal. The effects of p H and shear rate on beneficiation were not investigated. Only the effect of dosage was tested and is shown in Figures 6.3.3.3-1 and 6.3.3.3-2. The results indicate that the two non-ionic oil emulsions provide better separation than the cationic emulsion of kerosene, but worse than the tested anionic emulsion. combustible recovery 120 100 80^ o > o u 60£ 0) JQ . .to 40=J .a E o o 20 800 400 v 2 3 4 DOSAGE(%) Figure 6.3.3.3-1 Oil agglomeration of ultrafine Ford-4 coal using a kero-CO-610 emulsion (Ford-4 ultrafine coal, pH 7.7, 20,000 rpm, SHMP 100 ppm) Gaudin Index combustible recovery ash content 2 3 4 5 DOSAGE (%) 120 100 80 ;^ > 8 6o£ £ 40l E o 20° 200 800 40CP o JO Figure 6.3.3.3-2 Oil agglomeration of ultrafine Ford-4 coal using a kero-CO520 emulsion (Ford-4 ultrafine coal, pH 7.7, 20,000 rpm, SHMP 100 ppm) 151 6.3.3.4 Comparison of Beneficiation Results using Kerosene Emulsified with Different Surfactants The effects of various agglomerants on the beneficiation of Ford-4 ultrafine coal in a one-stage oil agglomeration are compared in Figure 6.3.3.3-4. As seen from the figure, for a 5% ash product, around 30% kerosene is needed in the conventional oil agglomeration process. A s Figure 6.3.3.3-4 shows, to obtain a product containing 8% ash, the required dosage is about 14% kerosene, 800 g/t of UBC-1 latex, or only 0.25% emulsified kerosene. In a one-stage separation, a product with 5% ash could be easily obtained at over 99.5%o coal recovery when 0.5%) kerosene emulsified with dodecyl sulfate is utilized. Although a product with 5% ash can also be produced at roughly the same recovery in the conventional oil agglomeration process, the required oil consumption (about 30%) is 60 times higher in comparison to the oil consumption when the anionic emulsion of kerosene is utilized. As seen from Figure 6.3.3.3-4, when oil addition is reduced to 0.5%, it is impossible to obtain a product with less than 15% ash at any coal recovery. Even when the oil dosage is increased to 5%, a product containing ash less than 12% can not be obtained, and is accompanied by a lower combustible recovery of approximately 92%. The tests with the use of the emulsified oil also demonstrated that the agglomerates separated on a 400 mesh screen were strong enough to withstand mechanical manipulation during processing. 152 20 < 200 400 600 800 (g/t) only for selective flocculation using UBC-1 j — i UBC-1 100 98 co g 96 o CD CH 0) V> 94 -O E o O 92 90 Kero-DDA Emulsion !<ero-DD^5 Emulsion Kerosene OnT F-4 ultrafine coal, natural pH7.8, 20,000 rpm 10 15 Oil Dosage (%) 20 (g/t) only for selective flocculation using UBC-1 2P0 4,00 6p0 800 25 • • Kero-DDS Emulsion Kero-DDA Emulsion Kerosene Only F-4 ultrafine coal, natural pH7.8, 20,000 rpm 120 15 10 10 15 20 Oil Dosage (%) 25 30 100 98 96 94 92 30 90 F i g u r e 6.3.3.3-4 C o m p a r i s o n o f o i l a g g l o m e r a t i o n u s i n g e m u l s i f i e d o i l w i t h c o n v e n t i o n a l o i l a g g l o m e r a t i o n a n d selective h y d r o p h o b i c a g g l o m e r a t i o n u s i n g the s a m e ultrafine c o a l (Test conditions: Ford-4 ultrafine raw coal, pH 7-8, oil agglomeration stirring rate 20,000 rpm, flocculation stirring rate 300 rpm, SHMP in agglomeration 100 ppm, SHMP in selective flocculation 300 ppm) 153 6.3.4 Summary and Discussion The emulsification and electrokinetic tests reveal that the surfactants reduce the size of the kerosene droplets and alter the electrical charge of the oil droplets. Regardless of the type of surfactants used, the size of the oil droplets can be reduced to a certain value with increasing surfactant concentration. Electrokinetic measurements with the emulsified kerosene show the strong effect of the added surfactants on the surface charge of the oil droplets. Dodecyl amine and C , 6 . , 8 long chain amine make the kerosene droplets either positively or negatively charged, depending on the p H and concentration of the amines. With an increase in amine concentration, the i.e.p. of the droplets can move from its original p H of 4.7 (kerosene only) up to around p H 11, which is close to the i.e.p. of the colloidal free amine (Laskowski et al. 1989). The experiments conducted with the strong electrolyte surfactants sodium dodecyl sulfate and sodium cetyl sulfate yielded a negatively charged droplets over the whole p H range. Significant differences in the abstraction behavior between the kero-DDA and kero-DDS droplets by coal are observed. The kinetic abstraction tests (Figures 6.3.2-1 and 6.3.2-2) indicate that the 200 mg of kero-DDA emulsion is entirely abstracted by coal in about 20 minutes, but only about 16% of the utilized kero-DDS emulsion is abstracted under the same conditions. The abstraction curves also indicate that the amount of kero-D D A abstracted by coal is much higher than that of kero-DDS. These differences of abstraction behavior may be related to the characteristics of the surfactants used and the electric charge of the coal particles. 154 It is well established (Capes & Germain 1989, Ignasiak et al. 1990, Keller & Burry 1990), and the results reported here confirm, that a relatively high oil consumption is required in conventional oil agglomeration. However, the tests carried out in this project have shown that much lower doses are sufficient when an emulsified oil is used. To control the dispersion of the mineral particles and to reduce the entrapment of gangue in the agglomerates, dispersants are utilized, and surfactants are employed to reduce the size of the oil-droplets and to control the electrical charge of the droplets. Under high shear hydrodynamic conditions, these well-dispersed oil droplets rapidly attach to the coal particles and thereby form agglomerates. The emulsions make the process more effective because the large surface area of the oil droplets offered increased collision opportunities with coal particles. Hence, emulsification allows oil to be more efficiently utilized, and makes it possible to reduce oil consumption. With the use of the kero-DDS emulsion, 0.5% or 0.25% oil addition was found to be sufficient to agglomerate ultrafine coal. At such low dosages, a clean coal product with an ash content around 5% can be obtained by screening, with over 99.9% combustible recovery. If a conventional oil agglomeration process is used instead, a much higher oil consumption is required (Figure 6.3.3.3-4). The role of surfactants in oil emulsification is dealt with in many reports (Shaw 1980, Shinoda & Friberg 1986, Myer 1988, and Laskowski 1992). The adsorption of surfactant at an oil/water interface decreases interfacial tension and reduces the work required to emulsify oil in water (Shaw, 1980). Imparting an electric charge to the oil droplets is another important aspect. The repulsion between the electrically charged oil droplets is an important factor stabilizing an o/w emulsion (Shaw, 1980). 155 A s is well known, surfactants are often utilized as collectors in froth flotation; water-insoluble oils and water-soluble surfactants-frothers are often used in coal flotation. Some surfactants may be employed as emulsifiers. In the emulsion flotation of molybdenite, sulphonate coconut oil is used as an emulsifier of oil (Hoover et al. 1976). When the oil, insoluble in water, is used as a collector in coal flotation, high intensity conditioning must be applied to disperse the oil into fine droplets. These droplets then collide and attach to hydrophobic particles. Therefore, pre-emulsification of oil followed by conditioning the pulp in fine coal flotation improves flotation (Sun et al. 1955, Bensley et al. 1977, Tyurnikova et al. 1981, Y u & Miller 1990, Laskowski 1992). In this project, all the tested oils were emulsified prior to use. During interaction between the oil droplets and coal particles, attachment of the oil droplets to coal particles is an important first step leading to oil agglomeration. The free energy change of the system during the attachment of an oil droplet to a solid particle is given by: A G = y o w ( c o s 0 - 1 ) (6.3.4-1) where A G is the free energy change of the system before and after the attachment of the oil droplet to the solid, y o w is the oil/water interfacial tension, and 0 is the three phase contact angle measured through water. Equation (6.3.4-1) shows that the attachment of an oil droplet to a solid particle may occur only when 0 > 0. Although no results in this study directly show the positive effect of surfactants in the oil-solid attachment, Schulman and Leja's study (1954) on the wettability of barite in the presence of 156 surfactants indicates that the surfactant dissolved in oil promotes the attachment of oil droplets to a solid surface. Schulman and Leja reported that the contact angle measured at oil droplets was much larger than at air bubbles. It was explained that the addition of a suitable surfactant could easily reduce the oil/water interfacial tension to a value of 1 mJ/cm 2 , but it was difficult to reduce the air/water surface tension to values below 30 mJ/cm 2. As explained by Melik-Gaykazian et al. (1967), the beneficial effect of surfactant (frother) on flotation with an oily collector results from its adsorption not only at the oil/water interface, leading to a lowering of the interfacial tension and improvement of the emulsification, but also at the coal/water interface to provide anchorage for the attachment of the oil droplets to the coal surface. Dowell M-210 promoter (a surfactant containing imidazoline) was used by Laskowski in a flotation study (1986) as a promoter. It reduced the kerosene/water interfacial tension from 50 mJ/m 2 down to 1 mJ/m 2 , and improved the spreading of kerosene on the coal surface. Brown et al. (1958) used m-cresol and dramatically improved the spreading coefficient of paraffin on coal. Y u et al. (1990) tested the effect of surfactants on the spreading of oil at the coal/water interface. They observed that as soon as the surfactant-emulsified oil droplets contacted the coal surface, the droplets immediately spread at the coal/water interface to form a thin oil "film" coating. Without the addition of surfactants, oil droplets formed a discrete thick patch on the coal/water interface after they were placed on the coal surface. This indicates that the surfactant enhances oil spreading at the coal/water interface. Accordingly, the coal recovery in the flotation tests with the use of the emulsified oil is always much higher in comparison with the results obtained with the unemulsified oil. 157 Oil a) Agglomerate formed in conventional oil agglomeration water b) Agglomerate formed in oil agglomeration using emulsified oils Figure 6.3.4-1 Schematic comparison of agglomeration using emulsified oils with conventional oil agglomeration As is well established, oil agglomeration requires intensive agitation to disperse the oil into small oil droplets in order to create opportunities for collisions between the oil droplets and coal particles. The oil droplets must attach to the coal particles first, the particles carrying oil then collide with each other, and finally form agglomerates via the bridging oil. Based on the observations made during the course of this project, a large amount of oil is required in a conventional oil agglomeration process (Figure 6.3.4-l(a)). In this case, the typical top size of the coal particles is about 200 mesh (74 um) (Capes 1982). Based on the size measurement of emulsified oil droplets (without surfactants) produced in a blender at 20,000 rpm, the top size of the oil droplets is about 100 um and the mean size (d so) is about 20 pm. The ratio of the top coal particle size to the oil droplet size is 74/100 (=0.74). In other words, the size of the oil droplets is almost the same as, 158 or coarser than, the size of the coal particles. Due to the relatively large size of the oil droplets, thick oil lenses are created between the coal particles, although the oil droplet may to some extent spread on the coal at such high intensity stirring. During the agglomeration process, the formed agglomerates may be broken by the vigorous shearing forces and reform again, consequently the agglomerates may be compacted to a certain degree. Nevertheless, conventional oil agglomeration consumes a significant amount of oil. However, when emulsified oil is used, the agglomeration process can benefit from the addition of surfactants. In the case of the agglomeration of ultrafine coal by means of oil emulsified with a surfactant, the top size of the ultrafine coal is 45 pm and the mean size (d50) is about 10 pm. The mean size (d50) of the emulsified oil droplets, which are produced in a blender at 20,000 rpm with the addition of a surfactant (e.g. DDS), is about 1.5 u,m. Thus, the ratio of the mean size of the coal particles to that of the oil droplets is about 7. In other words, the size of the emulsified oil-droplets is much smaller than the size of the coal particles. The numerous fine oil droplets dramatically increase the probability of collision between the coal particles and oil droplets. In addition, the surfactants improve the spreading of the oil droplets on the coal particles. As shown in Figure 6.3.4-1 (b), small and discrete oil droplets and relatively thin oil patches form on the Coal surface. When the coal particles carrying fine oil droplets collide with each other, small oil contact lenses will form between the coal particles. Accordingly, compact and minimum-oil-filled agglomerates are produced. This may explain why the required amount of bridging oil is much lower when the oil is emulsified. 159 A s shown in the oil agglomeration tests, both kero-DDA cationic and kero-DDS anionic emulsions effectively beneficiate ultrafine Ford-4 coal (Figures 6.3.3.1-1 to 6.3.3.1-3 and Figures 6.3.3.2-1 to 6.3.3.2-3) in spite of their completely different electrical properties. However, these results show that better selectivity is obtained when kero-DDS is used. This may result from the fact that most minerals at p H 7-8 carry a negative charge and can interact better with positively charged oil droplets. 100 50 0 -50 -100 40 20 0 -20 -40 -501 ) — e- „_. • s\ • 10(-2) M KCI - KERO DDA V - K E R O - DDS m • 1 2 4 6 8 10 12 — - — ^ - F O R D I N G - 4 IN 10(-2) M KCI - ^ ^ F O R D I N G - 1 3 \ ^ -S<vSILICA ' _ • • 2000 1500 1000 500 0 100 85 70 55 40 25 10 KERO-DDA KERO-DDS T5 TT O KERO-DDA . KERO-DDS 10 11 12 9 10 11 100 85 g 70 Q I-55 uj 4 0 g 25 « < 10 Figure 6.3.4-2 Zeta-potentials of the emulsified oil droplets and solid particles and corresonding oil agglomeration results 160 Figure 6.3.4-2 shows the zeta-potentials of the solid particles and the oil droplets and the oil agglomeration results. In the case of the kero-DDA, in an acidic environment the cationic oil droplets agglomerate coal particles in spite of electrostatic repulsion forces between the coal particles and the oil droplets. Recoveries in acidic solutions are therefore very high. The fact that the cationic emulsion agglomerates positively charged coal particles very well suggests that there are highly attractive structured forces which can overcome the electric repulsion. As demonstrated by X u and Yoon (1989, 1990) and Yoon et al. (1996), this hydrophobic attractive force is much larger than the dispersion force considered in the D L V O theory. As the figure shows, a large amount of silica is also entrapped in the agglomerates, which causes a low selectivity (low Gaudin Index) and low ash rejection. This may result from the strong electrostatic attraction between kero-DDA droplets and silica particles. When the p H approaches neutral and slightly alkaline values, both coal and silica are negatively charged, and there exists electric attraction forces between these particles and the kero-DDA droplets. According to the D L V O theory, kero-DDA should preferentially attach to silica due to the strong electrostatic attraction. However, the kero-DDA droplets preferentially attach to coal rather than silica because of the strong hydrophobic interaction with the hydrophobic coal. Therefore, as the figure shows, agglomeration selectivity (Gaudin Index) and ash rejection are substantially improved at around p H 8. Under alkaline conditions, the ash rejection still remains high but the recovery drops due to the electrical repulsive forces. As Figure 6.3.4-2 shows, kero-DDS is negatively charged over the whole p H range. At pHs below 7, both the electrostatic attraction between kero-DDS and coal, and the electrostatic repulsion between kero-DDS and silica, facilitate preferential 161 agglomeration of the coal particles by the kero-DDS. However, the hydrophobic interaction forces between the oil droplets and coal still prevail during agglomeration in an alkaline environment. Hydrophilic silica is not agglomerated by oil droplets irrespective of the electrical interactions. However, silica can be agglomerated by the kero-DDA. This is a reason why the kero-DDS is more selective. This fact further confirms that there are hydrophobic interaction forces between the oil-droplets and coal that play a very important role in beneficiation, and that hydrophilic particles are not agglomerated by the oil droplets. Although hydrophobic forces dominate the overall interactions between these solid particles and oil droplets during agglomeration, the electrostatic forces between the oil droplets and both solids render anionic kero-DDS more selective than cationic kero-DDA over the whole pH range. As described before, the hydrodynamic conditions in oil agglomeration are quite different from those in flocculation. Vigorous conditioning, which is utilized in agglomeration, provides very high kinetic energy collisions between coal and oil-droplets. These initial forces may be large enough to overcome any electrostatic repulsion and bring the oil-droplets and coal particles close enough to the distance at which attraction forces prevail (Yoon et al. 1996). Therefore, irrespective of the electrical charge of the oil-droplets, they are able to collide with coal and other solid particles under such hydrodynamic conditions. However, not all collisions lead to attachment. Whether the collision is successful or not depends on the hydrophobicity of the solid particles. The more hydrophobic the coal surface, the higher the probability that the attachment will be successful. Yoon et al. (1996) observed strong attractive forces when the contact angle 9 >90°. 162 6.4 Dewatering of Ultrafine Coal 6.4.1 Parameters Influencing Filtration of Ultrafine Coal There are a number of parameters which affect the filtration process, including particle size distribution, particle surface wettability, solid content, pH of the slurry, temperature, filtration time, slurry pretreatment, and hydrodynamic conditions, as well as filter driving forces that are related to the filtration equipment used. Among them, the most important elements are coal particle size, coal surface wettability, and slurry pretreatment. Although the effects of particle size and surface wettability on filtration are well recognized, studies on the dewatering of ultrafine (-45 pm) and fine (-500 pm) coal with the addition of chemical additives have not yielded much quantitative data, and did not explain the differences between the filtration behavior of ultrafine and fine coal. In the previous sections, chemical reagents such as flocculants, latices and oil emulsions, have been tested for the beneficiation of ultrafine coal. Since beneficiation is followed by dewatering of the products, the effect of such chemical reagents on filtration is also studied in this dissertation. 163 6.4.2 Filtration with the Use of Flocculants Including Poly electrolytes, Semi-Hydrophobic Flocculants and Hydrophobic Latex A series of tests were performed to evaluate the effect of flocculant dosage, hydrodynamic conditions and slurry p H on the filtration of the three types of ultrafine coal. The influence of three different additives on the filtration rate and cake moisture content was investigated under varying process conditions. 6.4.2.1 Effect of polyelectrolytes and hydrophobic latices on coal surface wettability Since the wettability of the coal particles is one of the most important parameters affecting filtration, the effect of the tested additives on the wettability of coal has also been studied. Receding contact angles on a polished coal surface were measured as a function of the concentration of the polyacrylamide flocculant, as well as the U B C - 1 and F R - 7 A latices (Figure 6.4.2.1-1). As can be seen from the figure, the measured receding contact angles indicate that the latices (UBC-1 and FR-7A) make the surface more hydrophobic while the polyelectrolyte (polyacrylamide) makes the coal surface more hydrophilic. This clearly demonstrates the fundamental differences between polyelectrolytes and totally hydrophobic latices. 164 6.4.2.2 Filtration of hydrophobic metallurgical coal and semi-hydrophobic coal using the three different types of flocculants Figures 6.4.2.2-1 and 6.4.2.2-2 show the results of fdtration tests with ultrafine Ford-4 coal and three flocculants. The flocculation tests reveal that the Ford-4 coal can be flocculated by all of the tested flocculants (Figures 6.1.2-1 to 6.1.2-3). The size of the floes appears to be largest when FR-7A or UBC-1 is utilized and smallest with PEO at the same dosage. The best filtration rate is observed when FR-7A is used (since not all tests were carried out with the use of FR-7A and UBC-1 it is stated that the best filtration rates were obtained with FR-7A but it is likely that this applies to both latices). However, as Figure 6.4.2.2-2 reveals, under these conditions, the cake moisture content is also very high. A similar trend is observed in the filtration of the semi-hydrophobic LC-7 coal (Figures 6.4.2.2-3 and 6.4.2.2-4). The filtration rate and moisture content in the filtration tests with LC-7 are lower than in the tests with the Ford-4 coal. However, the filtration rate is also much higher when hydrophobic latex (FR-7A) is utilized. As these results indicate, the cake moisture content increases whenever the filtration rate improves. 165 80 -| (Thousands) Dosage (g/t) Figure 6.4.2.2-1 The effects of flocculant type and dosage on filtration rate of Ford-4 ultrafine coal 45 -44 -43 -FR-7A , | Ford-41 42 - PAM ""S. itent ( 41 - • ture cor 40 - 1/+ f +^*» PEO Mois 39 -38 -37 -36 --t-ASH 9.10% 300 rpm -710 mmHg NATURAL pH 35 -C 0.2 0.4 0.6 (Thousands) 0.8 1 Dosage (g/t) Figure 6.4.2.2-2 The effects of the flocculant type and dosage on cake moisture content of ultrafine Ford-4 coal 166 It is well documented that hydrodynamic conditions have a strong effect on flocculation (Water & Loo 1984, Lynch & Hovak 1991, Hogg et al. 1993, Warren 1992). Figures 6.4.2.2-5 to 6.4.2.2-8 show the influence of shear rate on fdtration rate and cake moisture content for the hydrophobic Ford-4 and semi-hydrophobic L C - 7 coals at p H 6.4, and at a flocculant dosage of 100 g/t. As shown in these figures for hydrophobic Ford-4 coal, the filtration rates reach a maximum value at a stirring rate of 300 rpm when using P E O or P A M , and at 100 rpm when using FR-7A. For the semi-hydrophobic coal, however, both the filtration rate and cake moisture reach the highest levels at a stirring rate of 200-300 rpm. Such behavior might result from large but very loose floes at low shear rates and small but compact floes at higher shear rates. The maxima on the filtration rate curves correspond well with the maxima on the cake moisture content versus shear rate curves (Figures 6.4.2.2-5 vs. 6.4.2.2-6, Figures 6.4.2.2-7 vs. 6.4.2.2-8). Figures 6.4.2.2-9 and 6.4.2.2-10 show the effect of slurry p H on the filtration rate and cake moisture content of the hydrophobic Ford-4 coal. The results indicate that, although the best filtration rate is achieved at a neutral p H range of 6 - 8, the influence of p H is not overly critical. 20 0 100 200 300 400 500 600 700 800 D O S A G E (g/t) Figure 6.4.2.2-3 The effects of flocculant type and dosage on filtration rate of ultrafine LC-7 coal 36 0 100 200 300 400 500 600 700 800 D O S A G E (g/t) Figure 6.4.2.2-4 The effects of flocculant type and dosage on cake moisture content of ultrafine LC-7 coal 80 100 200 300 400 5oTT STIRRING RATE (rpm) Figure 6.4.2.2-5 The effects of stirring rate and flocculant type on fdtration rate of hydrophobic Ford-4 ultrafine coal 45 39 I -2g ^^^^^^^^^^^^^^^^^^^^^^^^^^^^^^^^^^^^^^^^^^^^^^^^^^^^^^^^^^^^^m 100 200 300 400 500 STIRRING RATE (rpm) Figure 6.4.2.2-6 The effects of stirring rate and flocculant type on cake moisture content of hydrophobic Ford-4 ultrafine coal Figure 6.4.2.2-7 The effects of stirring rate and flocculant type on filtration rate of ultrafine LC-7 Coal 200 300 400 500 STIRRING RATE (rpm) Figure 6.4.2.2-8 The effects of stirring rate and flocculant type on cake moisture content of ultrafine LC-7 coal I I LU Ford-4 DOSE 100 G/T 300 rpm -710mmHg pH Figure 6.4.2.2-9 The effect of pH on filtration rate for ultrafine Ford-4 coal o o 111 ' 40 QL Z> t— Ford-4 DOSE 100 G/T 300 rpm -710mmHg pH Figure 6.4.2.2-10 The effect of pH on cake moisture for ultrafine Ford-4 coal 171 6.4.2.3 Filtration of hydrophilic oxidized coal with the three different types of additives As previously shown, the use of FR-7A totally hydrophobic latex improved filtration rate. However, it also increased the cake moisture content in the tests with hydrophobic and semi-hydrophobic coals. The hydrophilic Ford-13 coal behaved quite differently from the other two tested coals. As described in the flocculation section, only PAM could flocculate the hydrophilic Ford-13 coal. Accordingly, the filtration rate for Ford-13 increased significantly with the addition of PAM, while it was very low and almost constant when FR-7A, UBC-1 or PEO was applied. Since Ford-13 was well flocculated by PAM, a relatively high filtration rate and a higher cake moisture content resulted. As Figures 6.4.2.3-1 and 6.4.2.3-2 show, FR-7A and PEO do not improve the filtration rate of Ford-13 coal. 800 DOSAGE (G/T) Figure 6.4.2.3-1 The effect of flocculant type and dosage on the filtration rate of Ford-13 ultrafine coal Figure 6.4.2.3-2 The effects of flocculant type and dosage on the cake moisture content of Ford-13 ultrafine coal Figure 6.4.2.3-3 The effects of stirring rate and flocculant type on the filtration rate of Ford-13 ultrafine coal 173 O o LU CC Z> H W O 44 43.5 43 42.5 42 41.5 41 40.5 40 39.5 Ford-13 PAM PEO FR-7A DOSAGE: 100 G/T pH:NATURAL VACUUM:710mmHg 100 200 300 400 500 Stirring Rate (rpm) 600 700 800 Figure 6.4.2.3-4 The effects of stirring rate and flocculant type on the cake moisture content of Ford-13 ultrafine coal Figures 6.4.2.3-3 and 6.4.2.3-4 show the effect of hydrodynamic conditions on the filtration rate and cake moisture content of Ford-13 coal. These results confirm that the hydrophobic latex F R - 7 A and semi-hydrophobic flocculant P E O cannot interact with the oxidized coal. As shown in the figures, the filtration rate and cake moisture do not respond to the changes in stirring rate when F R - 7 A and P E O is utilized. 6.4.3 Filtration with the Use of Surfactants and Emulsified Oils Besides the flocculants, surfactants and oily hydrocarbons are also often utilized as filtration aids. A detailed study on the application of emulsified oil in the filtration of fine coal, however, has never been conducted. The tests described in the preceding ? 174 chapter demonstrate that emulsified oil is an extremely promising type of agglomerant in fine coal beneficiation. 6.4.3.1 Adsorption of surfactants on coal surface The experimental methods for the measurement of surfactant adsorption on coal surfaces are described in Chapter 5. Based on literature concerning the kinetics of surfactant adsorption on coal (Keller et al. 1979, Gala 1982), the rates of adsorption for most surfactants on coal are fast, and 5 - 1 0 minutes is usually sufficient to reach an equilibrium. Therefore, in these tests, each sample was stirred for 15 minutes with surfactant in order to achieve equilibrium. The adsorption isotherms for the surfactants are determined by measuring the concentration of the surfactants in the solution before and after conditioning for 15 minutes. The calibration curves for dodecyl amine chloride and sodium dodecyl sulfate are shown in Figures 5.1.3-2 and 5.1.3-3 (Chapter 5); the electrical potential of the ion surfactant selective electrode is plotted versus dodecyl amine chloride concentration or sodium dodecyl sulfate concentration. Using the calibration curves, any concentration of D D A or DDS in a solution can be found based on the measurement of its electrode potential. The adsorption isotherms for dodecyl amine chloride (DDA) and sodium dodecyl sulfate (DDS) on Ford-4 coal are plotted in Figures 6.4.3.1-3 and 6.4.3.1-4, respectively. 0.008 DDA equilibrium concentration (m mol/l) Figure 6.4.3.1-3 Adsorption of dodecyl amine on ultrafine Ford-4 coal at pH 5.9 0.004 eg .§. 0.003 0.002 <" 0.001 0.001 0.01 0.1 DDS equilibrium concentration (m mol/l) Figure 6.4.3.1-4 Adsorption of dodecyl sulfate on ultrafine Ford-4 coal at pH 5.7 176 In both cases, three distinct regions of adsorption are apparent. In region I, the adsorption density of D D A on coal (Figure 6.4.3.1-3) slowly increases with its concentration, which may indicate that in this region the D D A surfactant ions adsorb individually onto the coal surface mainly through ion exchange and electrostatic attraction. As the numbers of adsorbed surfactant ions increase, a negatively charged coal surface is gradually neutralized and at the same time, hemi-micellization begins (Region II). In other words, in region II the hydrophobic attraction forces between hydrocarbon chains dominate and the shape of the adsorption isotherm becomes steeper. In region III, the increase of adsorption slows down probably due to Coulobic repulsions. If this is assumed to correspond to the monolayer coverage, it is 6.3 x 10"3 mmol/m2. This number can be used to calculate the area occupied by D D A molecules, resulting in 26.37 A 2 per molecule. This is different from the values of 10.75 A 2 for the area per adsorbed D D A molecule assumed by Castro et al. (1985) and the value of 16.4 A 2 calculated from the adsorption by Gala (1982), but it is fairly close to the value of 23 A 2 suggested by Arsentiev and Leja (1976). As Figure 6.4.3.1-4 shows, the adsorption isotherm for dodecyl sulfate is very similar to that of dodecyl amine, and again three adsorption regions are present. The initial adsorption of individual DDS ions onto a coal surface may result from electric attraction, chemisorption (Nicol, 1976), or/and hydrophobic attraction. These different adsorption forces will result in quite different initial orientations of the surfactant ions on the coal surface (Region I). Regardless of the reason leading to the initial adsorption, when the number of adsorbed surfactant ions is sufficient, hemi-micellisation apparently occurs (Region II). Assuming a monolayer adsorption takes place at the end of Region II 177 (about 3.1 x 10 3 m.mol/m2), this gives 53.57 A2 per molecule. Keller et al. (1979) calculated the surface area occupied by an adsorbed dodecyl sulfate ion to be 59 A 2 based on their adsorption isotherm for DDS on coal. Tajima et al. (1970) obtained 52 A 2 per molecule from their adsorption tests. Based on the above adsorption phenomena and analysis, the adsorption isotherms (Figures 6.4.3.1-3 and 6.4.3.1-4) indicate that both dodecyl amine and dodecyl sulfate adsorb onto coal and in both cases hemi-micellisation plays an important role. Wen and Sun (1977) showed through electrokinetic measurements that dodecyl amine adsorbs onto coal, quartz and pyrite. The adsorption predicted from the zeta-potential coefficient strongly points out a higher adsorption of the amine on coal than on the other two tested minerals. The high volatile bituminous coal, which was subjected to oxidization at 125 °C for 48 hours, could be floated with dodecylamine, which indicates that the amine makes the coal surface hydrophobic. In line with Brookes and Bethell's (1984) findings, the adsorption of amine increases the surface contact angle and improves the filtration rate, and at the same time filter cake moisture is reduced. Sobieraj and Majka-Myrcha (1980) also used electrokinetic experiments to study the interaction of dodecyl amine hydrochloride and sodium dodecyl sulfate with three different coals and the effect of coal oxidation on adsorption. Their results show quite convincingly that both D D A and DDS adsorb onto coals irrespective of coal oxidation. Their results seem to indicate that while DDS adsorbs onto the tested coals mainly in acidic solutions, the adsorption of D D A was obvious in all pH ranges. This correlates very well with the electrokinetic measurements by Woodburn et al. (1988), which clearly indicate adsorption of DDS onto coal over an 178 acidic and neutral p H range, and adsorption of the cationic surfactant, decyl trimethylammonium bromide, in the whole p H range. Table 6.4.3.1-1. Aggregation of ultrafine coal using D D A and D D S (natural p H 6.4) Initial cone, (mg/l) Dodecyl amine chloride Sodium dodecyl sulfate 0 Turbid and stable Turbid and stable 0.9 Turbid and stable Turbid and stable 1.8 Turbid and stable Turbid and stable 3.6 Turbid and stable Turbid and stable 7.2 Turbid, stability is dropping Turbid, improved stability 18 Apparent small aggregates (0.1-0.3 mm), settle down to the bottom in 4 minutes Turbid, improved stability 36 Apparent small aggregates (0.1-0.4 mm), settle down to the bottom in 1.5 minutes, very clear supernatant Turbid, improved stability 72 Apparent aggregates grow up (0.2-0.5 mm), settle down to the bottom in 1 minute, clear supernatant Turbid, improved stability 100 The aggregates are getting bigger, settle down to the bottom in 1 minute, clear supernatant Turbid, improved stability 144 The same as above. Turbid, improved stability These quoted conclusions can now be compared with the aggregation tests with the ultrafine Ford-4 coal (Table 6.4.3.1-1). In these tests 18 g of coal was first dispersed in a blender at 20,000 rpm for 1 minute, and then the slurry was transferred into a beaker with four baffles and continually stirred at 350 rpm for 10 minutes at a natural p H (6.4). 179 Following the addition of the predetermined amount of the surfactant, the slurry was stirred for 5 minutes more and then its stability was observed visually. As seen from the table, at a given concentration of dodecyl amine, the slurry becomes unstable, and the aggregates appear and then settle down. At higher D D A concentrations this results in a clear supernatant. In the case of DDS, the suspensions become quite stable at higher concentrations. The aggregation observed in suspensions of the ultrafine Ford-4 coal in the presence of D D A indicates that this surfactant decreases the value of the zeta-potential of the coal particles and this reduces the Coulombic repulsion forces between particles. A l l those who study flotation of oxidized coals observe that such coals can be floated with DDA. This suggests oriented adsorption of D D A ions onto coal with the hydrocarbon radicals facing the aqueous phase. However, this only suggests the overall result because coal is heterogeneous, and at the molecular level the orientation at various micro-areas can be quite different. Whereas the increased stability observed in the tested coal suspensions in the presence of DDS confirms adsorption of this surfactant onto coal, it also indicates an increased value of the zeta-potential which is consistent with the data published by others. 180 6.4.3.2 Filtration of ultrafine hydrophobic coal with surfactants and emulsified oils The filtration of ultrafine Ford-4 coal enhanced by surfactants and emulsified oils is shown in Figures 6.4.3.2-1 and 6.4.3.2-2. Since this coal can be well flocculated by hydrophobic latices such as F R - 7 A and U B C - 1 , its filtration rate is drastically improved. Large floes and high filtration rates are also observed using emulsified oils (kerosene emulsified with dodecyl amine or sodium dodecyl sulfate). At higher dosages, the filtration rate is improved by kero-DDA more than by any tested hydrophobic latex. The increase in filtration rate with the addition of P A M and surfactants (DDA) is insignificant. A s shown in the filtration tests with flocculants and ultrafine coal, whenever the filtration rate is improved, the cake moisture simultaneously increases as well. Based on experimental observations, this results from the large and strong floes that have trapped a large amount of water. With the cationic kero-DDA emulsion, even larger floes and clear supernatant are produced. It is not surprising that a high moisture content finally results. The cationic surfactant (DDA) has no effect on the cake moisture content. Anionic oil emulsion (kero-DDS), however, enhances the filtration rate and simultaneously reduces the cake moisture content of the ultrafine coal. The filter cake moisture with anionic surfactant (DDS) can be reduced by about 8% after a lengthy filtration, but filtration rate has to be sacrificed and is found to be very low due to the presence of a stable slurry without any aggregates. 100 90 80 70 60 50 40 I i 30£ 20' natural pH (-6.8), 300 rpm DDS "Kero-DDA <ero-DDS UBC-1 PAM 0 100 200 300 400 Flocculant Dosage (g/t) 500 600 0 1 2 3 4 Oil Emulsion Dosage (%) 5 6 0 400 800 1200 1600 2000 Surfactant Dosage (g/t) Figure 6.4.3.2-1 The effects of surfactants and emulsified oils on filtration rate of ultrafine Ford-4 coal Figure 6.4.3.2-2 The effects of surfactants and emulsified oils on cake moisture content of ultrafine Ford-4 coal 90 0 1 2 3 4 5 6 DOSAGE(%) 183 A s s h o w n i n F igure 6.4.3.2-1, the anionic k e r o - D D S emuls ion substantial ly improves the filtration rate, but its effectiveness declines w h e n it is over dosed. E x a c t l y the same trend is observed i n the f i l t rat ion o f the same ultrafine coal w i t h the addi t ion o f emuls i f i ed Esso heavy oi l s . F igure 6.4.3.2-3 shows the effects o f various emuls i f ied Esso o i l s o n f i l t ra t ion rate. A s seen f rom the figure, for any tested o i l emuls i f ied w i t h D D S the o p t i m u m dosage is a lways about 1%. The anionic o i l emulsions s ignif icant ly improve the filtration rate, and at the o p t i m u m dosage, also reduce the cake moisture content (Figure 6.4.3.2-4). In addi t ion, accord ing to Figures 6.4.3.2-3 and 6.4.3.2-4, the o p t i m u m dosages that produce the best filtration rate and the lowest cake moisture content are the same (about 1% dosage). U n l i k e the anionic emulsions, complete ly different behavior is observed i n the f i l t ra t ion o f the same ultrafine coal w i t h the cat ionic emuls ions . In this case, the enhancement o f filtration rate is direct ly proport ional to the increase i n cake moisture content (Figures 6.4.3.2-5 and 6.4.3.2-6). Th i s is i n l ine w i t h the effect o f the f locculants . F igures 6.4.3.2-7 and 6.4.3.2-8 show the effect o f slurry p H o n the filtration rate and cake moisture content o f hydrophobic Ford-4 ultrafine coal that was agglomerated w i t h emuls i f ied kerosene. W i t h the increase o f p H , in i t i a l ly the f i l t rat ion rate is s l igh t ly i m p r o v e d by a l l o f the cat ionic emulsions tested, but beyond a p H o f 9 or 10, the filtration rate drops drast ical ly. The slight influence o f p H o n the cake moisture content w i t h the use o f the cat ionic emuls ions is shown i n F igure 6.4.3.2-8. H o w e v e r , i n the presence o f anionic emuls ions , the filtration responds differently. B o t h the filtration rate and cake moisture content remain almost constant at ac id ic p H s but drast ical ly decrease i n an 184 alkaline environment. Obviously, the different fdtration behavior results from the nature of the interactions between the coal and the oil droplets as affected by the change of pH. 200 150 a l/mi E, —1 m Rate 100 ion M ^ — g . 2 LZ 50 0 M ESSOt156.904*1%DOAHCL Q ESSO1156H%0OAHCL \£ Kerosen+1%DDAHa Q ESS 0846* 1 %ODAHCI 0 ESSOB0S*1%DDAHa ^ ESSO2600*1%DDAHCI 0 1 2 3 4 5 6 Dosage (%) Figure 6.4.3.2-5 Filtration rate of Ford-4 ultrafine coal using cationic oil emulsions 50 Dosage (%) Figure 6.4.3.2-6 Cake moisture content of Ford-4 ultrafine using cationic oil emulsions 185 0) "5 o 70 60 50 40 30 DDA-1 DDA-5 O C16-18-4 ^ SCTS-1 n DDS 8 10 11 12 pH Figure 6.4.3.2-7 The effect of pH on filtration rate of Ford-4 ultrafine coal tested with kerosene emulsified with either cationic or anionic surfactants (DDA-1: kero emulsified in 5.4X10-4 M DDA solution, DDA-5: kero containing 1% DDA emulsified in distilled water, C16-18-4: kero containing 1% C16-18 amine emulsified in distilled water, SCTS-1: kero emulsified in 2.7x10-" M SCTS solution, DDS: kero emulsified in 3.4x10-4 M SDDS solution) 46 44 - 42 3 "E o O CO 40 38 36 DDA-1 Q DDA-5 0 C16-18-4 ^ SCTS-1 n DDS 7 PH 10 11 12 Figure 6.4.3.2-8 The effect of pH on cake moisture content of Ford-4 ultrafine coal tested with kerosene emulsified with either cationic or anionic surfactants 186 6.4.3.3. Filtration of hydrophilic ultrafine coal with surfactants and emulsified oils Unlike the filtration of the hydrophobic ultrafine coal described above, hydrophilic coal responds to the surfactants and emulsified oils quite differently. As depicted before, all of the tested surfactants (except DDS) and oil emulsions improved the filtration rate of the hydrophobic ultrafine coal to a certain extent. However, as can be seen from Figure 6.4.3.3-1, only the cationic kero-DDA emulsion has the ability to significantly increase the filtration rate of the Ford-13 ultrafine coal. The filtration rate is almost constant when the other reagents, including D D A , SDDS surfactants and kero-D D S anionic emulsion, are applied. Accordingly, as Figure 6.4.3.3-2 reveals, kero-DDA also increases cake moisture content due to the large and strong agglomerates that trap a significant amount of water. 1 5 0 3 ro a: c o K e r o - D D A P A M Kero-DDS 1 0 0 2 0 0 3 0 0 4 0 0 5 0 0 F l o c c u l a n t D o s a g e (g/t) N a t u r a l p H ( - 6 . 1 ) 3 0 0 r p m 7 1 0 m m H g D D A D D S ~FR=7A~ 6 0 0 -i s 1 s-O i l E m u l s i o n D o s a g e (%) 7 0 0 -1-T o T Sflo S u r f a c t a n t D o s a g e (g/t) fl 8 0 0 Sfio" ITJOTS 1 2 ^ TtuTJ TSbo Figure 6.4.3.3-1 The effects of surfactants and emulsified kerosene on filtration rate of hydrophilic Ford-13 ultrafine coal 187 o O P o 300 400 500 Flocculant Dosage (g/t) il Emulsion Dosage (%) Natural pH (-6.1) 300 rpm 710 mmHg 200 400 600 800 1000 Surfactant Dosage (g/t) 1200 800 T61 ho Figure 6 A3 3-2 The effects of surfactants and emulsified kerosene on cake moisture content of hydrophilic Ford-13 ultrafine coal Although the agglomerates also formed when the anionic kero-DDS emulsion was used, clear supernatant is not obtained since a certain amount of ultrafine particles with an extremely high negative charge are left dispersed in the solution. These dispersed ultrafine particles make filtration very difficult. Therefore, any improvement in the filtration rate of the hydrophilic coal is very limited with the anionic emulsion (kero-DDS). The filtration results with this hydrophilic coal using various emulsified oils are shown in Figures 6.4.3.3-3 and 6.4.3.3-4. In combination with the measurement of the 188 zeta-potential of the coal particles and oil droplets (Figures 4.1.1-1 and 6.3.4.2-1 to 6.3.4.2- 6), these results also demonstrate the complexity of the interactions in the studied systems. As seen from Figures 6.4.3.3-1 and 6.4.3.3-2, without chemical additives the filtration rate and cake moisture content of the hydrophilic ultrafine Ford-13 coal are about 17 ml/min and 41%, respectively. Figure 6.4.3.3-3 indicates that all the tested emulsions, whether cationic, anionic, or non-ionic, are able to enhance the filtration rate by 2 - 5 times, but only in an acidic p H range. Among these emulsions, the cationic emulsions are the most effective in enhancing the filtration rate due to the strong electrical attraction between the oil droplets and this coal. The tested non-ionic oil emulsion is least effective. However, in most cases, the best filtration rate still corresponds with a high moisture content of the filter cake, and as shown in Figure 6.4.3.3- 4, the highest cake moisture content is again produced by the cationic emulsions. The non-ionic emulsion produces the lowest cake moisture. At alkaline pHs, however, totally different filtration behavior occurs. Regardless of the type of emulsion employed, the filtration rate drops drastically to the level that is achieved in the absence of any chemical additives. In other words, none of the tested emulsions improves the filtration of the hydrophilic ultrafine coal in an alkaline solution. Figure 6.4.3.3-4 Cake moisture content of hydrophilic Ford-13 ultrafine coal with the use of kerosene emulsified with various surfactants 190 6.4.4 Effect of Particle Size on Filtration with Various Additives As mentioned previously, coal particle size is a very important parameter that has a major impact on the filtration process. The effect of particle size was further tested. The same additives, which were used in the filtration of the -45 pm ultrafine coal, are utilized in this section to investigate the differences in the filtration mechanisms between the -45 pm and -500 pm coal fractions. The Ford-4 and Ford-13 coals were used in these tests to represent hydrophobic and hydrophilic coals. 6.4.4.1 Filtration of -500 m hydrophobic coal with flocculants, surfactants and emulsified oils In the filtration of ultrafine coal preheated with various flotation reagents and filtration additives, it is found that an improvement in the filtration rate is always followed by an increase in the moisture content of the filter cake. In the case of the -500 pm Ford-4 coal, however, the totally hydrophobic latices, UBC-1 and FR-7A, not only increase the filtration rate, but also reduce the cake moisture content (Figures 6.4.4.1-1 and 6.4.4.1-2). Both the filtration rate and moisture content of the -500 pm coal can be improved with the use of P A M , but not as much as with the hydrophobic latices. The oil emulsions increase the filtration rate even more than the totally hydrophobic latices and they also significantly reduce the cake moisture. While cationic surfactant does not affect the filtration rate of the -500 pm Ford-4 coal, DDS reduces it. DDA, which has no effect on the cake moisture when the ultrafine fraction is filtered, enhances the dewatering of the -500 pm coal. 1000 0 100 200 300 Flocculant Dosage (g/t) 400 500 600 0 1 2 3 Oil Emulsion Dosage (%) 4 5 5 0 200 400 600 800 1 Surfactant Dosage (g/t) Figure 6.4.4.1-1 The effects of flocculants, surfactants and emulsified oils on filtration rate of the -0.5 mm hydrophobic Ford-4 coal 23 600 Flocculant Dosage (g/t) o 1 a 3 1 i & Oil Emulsion Dosage (%) 0 200 400 600 800 TTJbB r2u0 Surfactant Dosage (g/t) Figure 6.4.4.1-2 The effects of flocculants, surfactants and emulsified oils on cake moisture content of the -0.5 mm hydrophobic Ford-4 coal 192 6.4.4.2 Filtration of-500 m hydrophilic coal with surfactants and emulsified oils Figures 6.4.4.2-1 and 6.4.4.2-2 show the effects of various chemical additives on the filtration rate and cake moisture content of the -500 urn hydrophilic Ford-13 coal. As can be seen, the totally hydrophobic latex (FR-7A) does not interact with hydrophilic coal. They do not affect the filtration rate and cake moisture of the -500 u,m hydrophilic coal. In contrast, all the other tested additives are able to improve the filtration of the -500 pm hydrophilic coal to a certain extent. P A M substantially reduces the cake moisture content and improves the filtration rate. It is worthy of mention that among these additives, the kero-DDA cationic emulsion turns out to be the best filtration additive: it greatly enhances the filtration rate from 17 to 900 ml/min, and drastically lowers the cake moisture from 28% to 17%. There is no improvement in the filtration rate with the addition of the kero-DDS anionic emulsion although it significantly reduces the cake moisture. The cationic surfactant is better than the anionic surfactant. It does not, however, significantly improve the filtration of hydrophilic coal. 0 100 200 300 400 Flocculant Dosage (g/t) 500 600 700 0 1 2 3 4 Oil Emulsion Dosage (%) 5 6 7 0 200 400 600 800 1000 1200 1400 Surfactant Dosage (g/t) Figure 6.4.4.2-1 The effects of flocculants, surfactants and emulsified oils on filtration rate of -0.5 mm hydrophilic Ford-13 coal 193 o 5 15 200 FR-7A Kero-DDA Natural pH (-6.0), 300 rpm, 710 mmHg 0 100 200 300 400 Flocculant Dosage (g/t) 500 600 700 0 1 2 3 4 Oil Emulsion Dosage (%) 5 6 7 400 600 800 Surfactant Dosage (g/t) 1000 1200 1400 Figure 6.4.4.2-2 The effects of flocculants, surfactants and emulsified oils on the cake moisture content of the -0.5 mm hydrophilic Ford-13 coal 6.4.5 Filtration of Ultrafine Coal under High Pressure The filtration experiments carried out with ultrafine coal reveal that slurry pretreatment with totally hydrophobic latices or emulsified oils significantly improves the filtration rate. The cake moisture content, however, is simultaneously increased in most cases. The observations show that the above filtration phenomenon is always accompanied by large and strong floes, estimated to be up to 3 -4 mm in size. It is proposed that a large amount of water might be entrapped in these big and strong floes. While an applied vacuum reduces the moisture existing between the floes, it is not able to overcome the capillary pressure inside the floes to release the trapped water. This hypothesis was tested by studying the effect of filtration pressure on the cake moisture content with and without the presence of hydrophobic flocculant, FR-7A. Figure 6.4.5-1 indicates that the trapped water can be released from the large floes produced by F R - 7 A at a certain pressure, which depends on the dosage of the agglomerant. Under a certain pressure, the cake moisture content is lower than that in the absence of F R - 7 A . For 194 instance, at pressure above 4 kPa, water is released from the floes produced by FR-7A at a dosage of 200 g/t, whereas pressures exceeding 8.5 kPa are needed to squeeze water out of the floes produced by FR-7A at 400 g/t. Each test was repeated once, and the average of the two values was used in the plots. P r e s s u r e (kg/cmA2) Figure 6.4.5-1 The effect of filtration pressure on the cake moisture content of ultrafine Ford-4 coal with the addition of hydrophobic latex FR-7A 6.4.6 Summary and Discussion A dewatering process can be divided into a few stages. If water saturated particles form a bed through which drainage occurs, three conditions can arise (Harris et al. 1957, Grey 1958): (a) Free flow through the packed particles will occur i f the surface is completely covered with water. This will be mostly determined by the bed porosity which is controlled by the particle size distribution. (b) Drainage of water from the bed when the water/gas interface appears within the packed layer of particles, which happens when the applied pressure exceeds the negative capillary pressure at the air/water interface. The system, when displacement 195 proceeds, reaches a funicular state, in which a continuous network of water exists in equilibrium with air throughout the bed. This process is quite strongly affected not only by the diameter of the capillaries and the solid surface properties but also by the liquid's surface tension. (c) At more prolonged or increased suction, the continuous liquid film surrounding the particles breaks and the water forms lenses around more hydrophilic areas, and/or particle-particle contacts. This is the pendular state. It is obvious that the presence of water droplets in a porous bed under such conditions is mostly determined by the solid surface wettability. From this brief analysis, it is clear that the tested filtration additives and their functions can be classified as follows: i) flocculants and emulsions increase filter cake porosity by aggregating fine particles; ii) surfactants lower water surface tension and reduce the capillary retention forces; and iii) oily hydrocarbons and their emulsions make solids more hydrophobic and agglomerate particles. The effect of the three tested additives is discussed separately below for each additive. Flocculants From the filtration results for the ultrafine coal, it is obvious that filter cakes in such experiments are characterized by very high moisture contents which further increase with flocculation. There is no doubt that this is caused by the extremely fine size of the treated coal particles (below 45 pm). The cake moisture content always increases when the flocculant improves the filtration rate of the ultrafine coal. It is proposed that this may result from a considerable amount of water trapped in the large and strong floes. This water cannot be removed from the floes under vacuum filtration conditions. As shown in Figure 6.4.6-1 (a), in a filter bed formed by large and strong floes, tiny capillaries exist between the ultrafine particles. During the dewatering process, air passes through the 196 voids between the floes and pushes the inter-floc water out of the cake, but it cannot overcome the capillary pressure within the floe. This "short circuiting" phenomenon might be responsible for the tremendous filtration rate improvement and high water retention within the floes. High pressure filtration must be applied in order to break the floes and release the water from the floes. For all of the tested -45 pm ultrafine coals, P E O is unable to improve the filtration rate. The hydrophobic latices (UBC-1 and FR-7A) dramatically improve the filtration rate of hydrophobic bituminous coal (Ford-4), but are entirely inefficient for the hydrophilic Ford-13 sample due to the fact that the hydrophobic latices do not interact with the hydrophilic surface of Ford-13 coal. The polyelectrolyte (PAM) provides only a very small increase in the filtration rate for the Ford-4 and L C - 7 samples, but it is definitely the best in the tests involving the Ford-13 coal. Coal particle size has a large effect on filtration. The hydrophobic latices (UBC-1 and FR-7A) always increase the cake moisture content for the -45 pm Ford-4 ultrafine coal as the filtration rate improves (by up to 300%). However, the hydrophobic latices reduce the moisture content for the -500 pm fines and simultaneously increase the (a) Ultrafine coal flocculated or agglomerated by flocculants or oil emulsions (b) Fine coal flocculated or or agglomerated by flocculants or oil emulsions Figure 6.4.6-1. The structure of filter bed formed by ultrafine and fine coal with addition of flocculants or other chemical additives 197 filtration rate by 50%. This finding reveals the substantial differences in the filtration mechanisms of the -500 um fine and -45 urn ultrafine coal. During the dewatering of the -500 um fraction (Figure 6.4.6-1 (b)), air can penetrate the filter bed through intra-floc voids between and partially within the floes because of relatively large capillary dimensions and the collapse of some loosely structured floes during the process. Similarly to the filtration of the hydrophilic ultrafine coal, the hydrophobic latices also have no effect on the filtration of the hydrophilic Ford-13 fine coal (Figures 6.4.2.3-1, 6.4.2.3-2, 6.4.2.3-3 , 6.4.2.3-4, 6.4.4.2-1 and 6.4.4.2-2). Emulsions The trend observed with flocculants, namely an increase in cake moisture of ultrafine coal with flocculation, is also observed with the use of the emulsified kerosene. This phenomenon may be explained by the same mechanisms described above. As shown in the filtration tests, cationic and anionic emulsions have different effects on the filtration of ultrafine coal. Figures 6.4.3.2-1 and 6.4.3.3-2 show that the kero-DDA emulsion can improve the filtration of both the tested hydrophobic and hydrophilic ultrafine coals. It is reasonable to assume that the positively charged oil droplets are able to attach to the slightly negatively charged Ford-4 particles and highly negatively charged Ford-13 coal particles around a neutral pH. At the same dosages, filtration rates improve much more for the Ford-13 than for the Ford-4 coal. As described in the preceding sections, with the rise of kero-DDA concentration, the adsorption density on the coal surface continuously increases, and this makes the coal surface more hydrophobic. At the same time, the size of the aggregates formed are constantly enlarged, and might reach 1 to 8 mm. The filter cake formed from these agglomerates contains a 198 great deal of moisture. The anionic kero-DDS emulsion, however, provides very promising results; at low dosages this additive can significantly increase the filtration rate, while also lowering the cake moisture content. As found from the abstraction of kero-DDS oil droplets by hydrophobic coal, the abstracted amount is much less than in the case of kero-DDA, and the abstraction on the coal surface will reach a maximum at a certain concentration. The kero-DDS at this concentration can make the coal surface more hydrophobic, and leads to a high filtration rate and a low cake moisture content. At higher doses of the kero-DDS emulsion, the large and loose aggregates and a turbid supernatant result in a high cake moisture and a low filtration rate. It is noteworthy that both the cationic and anionic emulsions exhibit excellent filtration/dewatering abilities for the -500 pm fine coal. They provide the best filtration rate (Figures 6.4.4.1-1 and 6.4.4.2-1). O f the two emulsions, kero-DDA yields the better dewatering improvement. A s shown in Figure 6.4.4.1-1, filtration rates resulting from the addition of kero-DDA emulsion are almost double those obtained with hydrophobic latices. A similar situation can be seen in Figure 6.4.4.2-1 for the hydrophilic Ford-13 coal. The advantage of kero-DDA emulsion over the hydrophobic latices and polyelectrolyte is likely due to its positive surface charge, which favours the attachment of the emulsion droplets to coal particles and makes the coal surface more hydrophobic. In addition, due to the loose aggregates formed by -500 pm particles, kero-DDA also significantly reduces the cake moisture content under normal vacuum pressure. In contrast, with kero-DDS, a negative charge on both the emulsion droplets and coal particles hinders this attachment. This leads to relatively low filtration rates. The hydrophobic Ford-4 coal particles, in spite of their slightly negative 199 surface charge at acidic pHs, are well agglomerated by kero-DDS droplets. The hydrophobic interaction between coal organic matter and oil droplets is apparently strong enough to overcome the weak electrical repulsion between them. This is clearly observed during the agglomeration tests in which kero-DDS leaves a relatively clear supernatant after agglomerating the Ford-4 but the supernatant is turbid when the Ford-13 coal is agglomerated. Therefore, kero-DDS substantially improves the filtration rate of the -500 pm Ford-4 coal, but not the Ford-13 coal. Filtration studies under varying pHs with the hydrophobic and hydrophilic ultrafine coals and with the addition of various emulsions were carried out and these results are shown in Figures 6.4.3.2-7, 6.4.3.2-8, 6.4.3.3-3 and 6.4.3.3-4. In combination with the zeta-potential measurements of coal particles and emulsified oil-droplets (Figures 4.1.1-1 and 6.3.1.2-1 to 6.3.1.2-6), these results provide significant evidence that the mechanism of the attachment of the oil-droplets to the coal surface involves electrostatic and hydrophobic interactions. 150 . )K 5.4x10(-3) M DDA.HCI <>3.9x10(-4)M C16-18Amine Q Ford-4 Coal particles "I" 2.7x10(-4) Sodium cetyl sulfate . 4.3x10(-3) M Sodium dodecyl sulfate ^ V , Q Ford-13 Coal 50 1--50 + + + + ± -100 1 — i • . . i • . • i I 0 2 4 6 8 10 12 14 PH Figure 6.4.6-2 The zeta-potentials of Ford-4 coal particles and oil droplets emulsified with cationic and anionic surfactants 200 The zeta potential measurements of Ford-4 coal particles and kero-DDA and kero-D D S oil droplets are summarized in Figure 6.4.6-2. This figure may facilitate understanding of the filtration results shown in Figures 6.4.3.2-7 and 6.4.3.2-8. In an acidic environment, electric attraction and n o n - D L V O hydrophobic forces are apparently predominant in the interactions between the anionic oil droplets and the Ford-4 coal particles, and strong agglomerates are consequently produced. This results in a high filtration rate but also a high moisture content. Figure 6.4.3.2-7 shows good filtration with kero-DDS in the p H range of 3 to 5. For the cationic oil droplets, an electric repulsion between the coal particles and cationic oil droplets can obviously be expected. However, this repulsion is obviously overcome by the strong hydrophobic forces, and agglomerates are still well formed. This is reflected in the high filtration rate and high moisture content. In alkaline aqueous solutions, the filtration rate and moisture content with the anionic emulsions are all reduced, which may be explained by the strong electrostatic repulsion between the coal and the anionic droplets. However, why this repulsion can not be overcome by the hydrophobic interaction forces is still difficult to explain. The strong agglomeration of the coal with the cationic oil emulsions may be attributed to the electric attraction and hydrophobic forces. This subsequently leads to an increasing filtration rate and a slightly higher moisture content. However, when the slurry p H is approaching p H 11, close to the i.e.p. of a colloidal amine, the electric attraction between the coal and kero-DDA droplets disappears, producing weak and loose agglomerates. Therefore, a relatively low filtration rate is observed when the cationic emulsion is utilized over a very alkaline p H range. Surfactants Although it has been known that surfactants can improve the filtration of fine coal to some extent (Keller et al. 1979, Dolina et al. 1971, Ofori et al. 1989), in general they 201 are not very efficient filtration additives. It is claimed that they facilitate dewatering by lowering capillary pressure (Silverblatt and Dahlstrom 1954, Nicol 1976, Keller et al. 1979). However, ionic surfactants also adsorb onto coal, neutralize the coal surface charge, and may increase coal hydrophobicity. A l l these effects will influence particle agglomeration. The results of this dissertation show that the effect of the surfactants on filtration varies considerably. The results of the surfactant adsorption tests may well explain the filtration results for -500 pm coal particles, but not so well the results for -45 pm ultrafine coal. D D A provides some improvement in the filtration rate for ultrafine Ford-4 coal (Figure 6.4.3.2-1), but is inefficient in the filtration of -500 pm fine coal (Figure 6.4.4.1-1) because of the reduced aggregation power for the coarse particles. As described in the surfactant adsorption tests, D D A may make the coal surface more hydrophobic, which is in line with the contact angle measurements by Brooks and Brethell (1979), and the flotation results obtained by Wen and Sun (1977). The filtration tests using D D A show a considerable moisture reduction for -500 pm samples of Ford-4 and Ford-13 coals (Figures 6.4.4.1-2 and 6.4.4.2-2), but not for the - 40 pm fraction (Figures 6.4.3.2-2 and 6.4.3.3-2). However, an overly high dosage of D D A increases cake moisture, and this may be caused by reverse adsorption. Brooks and Bethel (1984) studied the effect of aliphatic amine on the dewatering of a -500 pm coal. Their findings are similar to the results of this study. A minimum cake moisture content appeared at about 400 g/t of amine, whereas in this study, the minimum is observed at about 100 g/t. At higher dosages, an increase in the cake moisture is observed in both cases. Their optimum surfactant dosage (400 g/t), at which the minimum cake moisture content occurs, however, does not correlate well with the dosage (4 kg/t) where the maximum contact angle is observed. Anionic surfactant (DDS) affects the filtration process in another way. Figures 6.4.3.2-2 and 6.4.4.1-2 reveal that D D S is able to reduce the cake moisture for the ultrafine Ford-4 coal but not for the -500 pm fraction. The results obtained with these 202 surfactants contrast with those published by Nicol (1976) who, using cetyl dodecylammonium bromide and SDDS in the dewatering of the -500 + 76 pm fraction of coal, found the anionic surfactant to be more suitable as a coal dewatering additive. The adsorption of anionic surfactants on negatively charged coal is apparently lower than that of cationic surfactants. Combined with the observation made during the aggregation tests with Ford-4 ultrafine coal, SDDS seems to be able to make the coal surface more hydrophilic. Additional wettability measurements of coal in the presence of SDDS are necessary in order to better understand the interactions between coal and the S D D S surfactant. At very low concentrations, when the adsorption density at the solid/solution interface is low, surface tension reduction may be the dominant driving force in the dewatering mechanism with anionic surfactants. The present results are in agreement with those by Dolina and Kamisnki (1971) and Ofori et al. (1989), who showed that the cationic surfactants performed well with all tested coals. It is difficult to single out a common rule for all surfactants due to the great variations in the experimental data. These variations may be caused by different experimental conditions and different coals employed by those researchers. The surfactants have to be evaluated individually and more experimental data are needed for various surfactants and different coals to determine statistically reliable trends. The observations made in this project may provide some clues to explain some of the filtration results. Improvement in the filtration of ultrafine coal by surfactants is not so significant. Even for a -500 pm fine coal, high dosages of surfactants (preferably cationic) have to be used to achieve both improvement of filtration rate and reduction of cake moisture. _ 203 C H A P T E R 7 CONCLUSIONS A N D R E C O M M E N D A T I O N S Beneficiation of Ultrafine Coal with a Hydrophobic Latex 1. In processes in which fine coal particles are aggregated with the use of appropriate additives, the coal surface properties are shown to be as important as the properties of the additives. 2. The abstraction of the U B C - 1 hydrophobic latex on the hydrophobic coal and silica indicate that the hydrophobic latex has a much higher affinity to coal than to hydrophilic mineral matter. The attachment of the hydrophobic latex to the hydrophobic coal depends on slurry pH. Over an acidic p H range, the abstraction of the latex is relatively high, resulting in a near total agglomeration of the hydrophobic coal particles. In alkaline solutions, the attachment of the latex to the hydrophobic coal particles markedly decreases due to the strong eletrostatic repulsion between the coal and the latex. The more hydrophilic the coal, the poorer hydrophobic agglomeration with the latex. 3. Hydrophobic latices such as U B C - 1 and F R - 7 A effectively agglomerate hydrophobic coal particles via hydrophobic interactions but are unable to flocculate hydrophilic particles of oxidized coal and silica. In contrast, P A M , a water soluble polyelectrolyte, flocculates all tested coal particles suspended in water regardless of their surface wettabilty. 204 4. Beneficiation tests carried out with ultrafine coal and mixtures of coal/silica and coal/clay using hydrophobic latices reveal that the hydrophobic latex (UBC-1) developed in this Department is extremely selective and flocculates only hydrophobic particles of bituminous coal. The obtained aggregates are strong enough to withstand mechanical manipulation and can be easily separated on a screen from the dispersed tailings. This new process is very similar to oil agglomeration; however, the reagent consumption in this process is about 100 times lower than the oil consumption required in the conventional oil agglomeration process. The results of the one-step beneficiation process show that at the optimum conditions (i.e. p H 7-8, a stirring speed of 300 rpm and a latex dosage of 300 - 400 g/t), a clean coal product with an ash content of 8.1% can be produced on a screen at over 95% combustible recovery from the ultrafine Ford-4 run-of-mine coal with an ash content of 17.3%, 5. The beneficiation results confirm the strong effect of p H on hydrophobic agglomeration. The optimum selectivity at high recoveries is always observed around a neutral p H range. Well formed agglomerates are usually obtained at low p H values but their ash contents are very high, while better selectivity at alkaline conditions is always accompanied by lower coal recoveries. 6. Hydrodynamic conditions have a significant effect on the hydrophobic agglomeration of ultrafine coal. Some level of mixing is necessary to initiate hydrophobic agglomeration with the totally hydrophobic latices and to achieve the best selectivity. When hydrophobic latices are used, the optimum stirring speed is about 300 rpm for the device used in this study. In addition, the results indicate that the solids content of the slurry and the addition of dispersants also affect separation efficiency. A solids content of 2% and a sodium hexmetaphosphate concentration of 300 mg/l provided the best results. 205 7. The beneficiation of ultrafine coal with the use of hydrophobic latices has only been tested in a one-stage process. If a super-clean coal is required, a mutiple-stage separation is recommended in order to further reduce the ash content of the clean product. O i l Agglomeration of Ultrafine Coal 1. The abstraction of the emulsified oil by coal particles is found to be different for cationic (kero-DDA) and anionic (kero-DDS) emulsions. The amount of oil emulsified with the cationic surfactant abstracted by coal continuously increases with increasing oil concentration and the abstraction curve appears to be almost linear in the range of tested emulsion concentrations. No plateau density was observed in this range. The abstraction density of the anionic emulsion by the coal rapidly reaches a plateau at a very low oil concentration of about 0.5%. The amount of the anionic emulsion abstracted is more than 50 times lower than that of the cationic emulsion. 2. The agglomeration tests carried out with different type of coals and emulsions showed an obvious effect of coal surface properties. The agglomeration results strongly depend on hydrophobic interactions and, to a certain extent, on electrostatic forces between the emulsified oils and coal particles. 3. The results of the beneficiation of ultrafine coal with emulsified kerosene are very promising. With the kero-DDS emulsion, a 0.25% to 0.5% oil consumption is found to be sufficient to agglomerate the ultrafine coal. From an ultrafine coal feed with an ash content of 17.3%, a clean coal product at over 99% coal recovery with an ash content of about 5% is obtained by screening after one-stage agglomeration. Furthermore, the 206 produced agglomerates are strong enough to withstand mechanical manipulation in the process and can be recovered by screening. 4. Electrokinetic measurements of the oil droplets obtained by emulsification with various surfactants indicate that the surfactants make the oil droplets either positively or negatively charged. Addition of surfactants sharply reduces the size of the oil droplets, improves the stability of the emulsions, and enhances attachment and spreading of oil-droplets on the coal surface. This leads to a more efficient oil agglomeration process and a substantial reduction in the oil consumption. The electrical charge and oil-droplet size distribution depend on the type and amount of surfactants used for emulsification. The electrokinetic measurements also explain why the cationic emulsions work well with oxidized coal. 5. A l l surfactants which enhance oil emulsification improve the oil agglomeration process. The overall beneficiation results are more selective with anionic emulsions, while hydrophilic oxidized coal can be agglomerated only with the cationic emulsion. 6. As in a conventional oil agglomeration process, oil dosage and intensity of conditioning critically affect the oil agglomeration when emulsions are utilized. The separation results show that the higher the stirring rate, the better the beneficiation. Dewatering of Ultrafine Coal 1. A l l of the hydrophobic latices and emulsified oils that improve the beneficiation of ultrafine coal enhance the subsequent filtration process. For instance, the hydrophobic latex U B C - 1 or F R - 7 A not only successfully beneficiates the run-of-mine Ford-4 207 ultrafine coal but also significantly improves its filtration rate. As shown in the oil agglomeration tests using emulsions, depending on the coal surface properties, the kero-D D S anionic emulsion is one of the best oil agglomerants for metallurgical coal. This emulsion is also the best dewatering additive since it substantially improves the filtration rate and dramatically reduces the cake moisture content in the subsequent dewatering process. For the Ford-13 oxidized coal, the cationic kero-DDA gives better results. 2. The cake moisture content for ultrafine coal is much higher than that for 0.5 mm coal. For the ultrafine coal (-40 pm), the filtration process depends much more on the size of the particles than on their surface properties. 3. The three tested types of flocculants (UBC-1 and F R - 7 A , P E O , and P A M ) exhibit different aggregation abilities for ultrafine coal. The hydrophobic latex, which well aggregates ultrafine hydrophobic coal, will also significantly increase its filtration rate. However, it does not interact with the hydrophilic coal. P E O is unable to enhance the filtration of the tested ultrafine coal. Polyacrylamide is able to flocculate all of the tested coal samples regardless of their surface wettability, but produces relatively small and loose floes, and subsequently forms a less permeable filter cake for the hydrophobic coal in comparison with those produced by the UBC-1 or F R - 7 A hydrophobic latices. 4. It is found that for the tested ultrafine coal the cake moisture content almost always increases when the filtration rate is improved following the addition of flocculants or other filtration additives. This probably results from the large amount of water trapped in the large floes. 5. Significant differences in filtration behavior are observed between ultrafine coal (-45 pm) and fine coal (-0.5 mm). Different filtration mechanisms for ultrafine and fine coals with the addition of the filtration additives have been proposed. 208 6. Oil emulsification not only improves oil agglomeration and significantly reduces oil consumption in the oil agglomeration process, but also produces very powerful and promising filtration/dewatering aids for both the ultrafine and fine fractions of hydrophobic and hydrophilic coals. By selecting either cationic or anionic emulsifiers, even low rank/oxidized coals can be efficiently agglomerated. 7. In comparison with the other tested dewatering additives, surfactants alone are not effective in improving filtration. The same surfactants may, however, be very effective when used to emulsify oily hydrocarbons and then employed in the form of emulsions. Recommendations The new hydrophobic agglomeration process for ultrafine coal, which involves the use of hydrophobic latices or emulsified oil, should be further studied for different types of coals using different surfactants. The lab experiments should be scaled-up and also include continuous beneficiation, separation and dewatering tests. The effect of abstraction of emulsified oil droplets and adsorption of surfactants on coal wettability were not studied in this dissertation. To better understand the abstraction and adsorption mechanisms, these experiments should be included in future research projects on oil agglomeration and filtration involving these reagents. Such tests should provide correlations between surfactant adsorption, contact angle, zeta potential, and filtration results, as well as correlations between emulsified oil droplet abstraction, contact angle, zeta potential, and beneficiation results. 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Wt. g, o <0.1 0 . 07 0 . 07 0 . 01 0 . 01 0 .12 0 . 12 • 0 . 1 - 0 . 2 0 .14 0 . 21 0 . 07 0 . 08 0.43 0 . 55 0 . 2 - 0 . 3 0 . 34 0.54 0.22 0 .30 ' 0 .62 1.17 0 . 3 - 0 . 4 0 .41 0 . 95 0 . 36 0 . 66 0 .80 1. 97 0 . 4 - 0 . 5 0 .47 1.42 0.43 1. 09 1. 05 3 . 02 0 . 5 - 0 . 6 0 . 61 2 . 03 0 . 50 1. 59 1.05 4 . 08 0 . 6 - 0 . 7 0 . 74 2 . 77 0 . 72 2 .31 1.24 5 .31 0 . 7 - 0 . 8 0 . 95 3 . 72 0 . 94 3 .25 1.36 6 . 67 0 . 8 - 0 . 9 1. 15 4 . 87 1.08 4 . 33 1.48 8 .15 0 . 9 - 1 . 0 1.22 6 . 08 1.29 5 . 62 1.61 9 . 76 1 . 0 - 0 . 5 1.69 7 . 77 1.80 7.42 1.85 11 . 62 1 . 5 - 2 . 0 2 .16 9. 93 2 .16 9 . 58 2 . 22 13 .84 2 . 0 - 2 . 5 2 . 57 12 . 50 2 . 52 12 .10 2 .47 16 . 31 2 . 5 - 3 . 0 3 . 04 15 . 54 2 . 88 14 . 97 2 . 78 19 . 09 3 . 0 - 3 . 5 3 .38 18 . 92 3 .24 18 .21 2 . 90 22 . 00 3 . 5 - 4 . 0 3 . 92 22 . 84 3 . 53 •21. 74 3 .40 25 .40 4 . 0 - 5 . 0 4 .59 27.43 3 .96 25 .69 3 . 77 29 . 17 5 . 0-6 . 0 5 . 07 32 . 50 4 .46 30 .15 4 .33 33 .49 6 . 0 - 7 . 0 5 .47 37 . 98 5 . 04 35 .19 4 .64 38 .13 7 . 0 - 8 . 0 5 . 74 43 . 72 5 . 32 40 . 51 4 . 94 43 . 07 8 . 0 - 9 . 0 6.22 49 .93 5 . 76 46 .27 5 .25 48 . 33 9 . 0 -11 .0 6.62 56 . 56 6 . 33 52.60 5 . 56 53 . 89 1 1 . 0 - 1 2 . 0 6 .69 63 .25 6 .69 59 .29 6 . 00 59 .89 12 . 0-14 . 0 6 . 76 70 . 00 6.98 66 .27 6 .12 66 . 00 14 . 0-16 . 0 6 . 62 76 . 62 7 .19 73 .46 6 .12 72 . 12 16 . 0 -20 .0 6 .42 83 . 04 7 .19 80 . 66 6 . 18 78 .30 2 0 . 0 - 2 5 . 0 5 .88 88.92 6 .47 87.13 6 .12 84 .42 2 5 . 0 - 2 9 . 0 5 .41 94 . 33 5 . 76 92 . 89 6 . 06 90 .48 2 9 . 0 - 3 7 . 0 3 .38 97 . 71 3 .96 . 96 . 84 5 .25 95 . 73 37 . 0 -43 .0 1. 35 99. 06 1.44 98 .28 3 .40 99 . 13 43 . 0 -45 .0 0 . 34 99 .39 0 . 36 98 . 64 0 . 62 99 . 75 >45 . 0 0 . 61 100.00 1.36 100.00 0 .25 100 .00 TOTAL 1 100.00 100.00 100 .00 227 iquid solid ICUL (a) v a c u u m v a c u u m T ICUL (b) T acu (c) v a c u u m Figure Apdx - 1. The bed configurations during filtration and dewatering stages Table Apdx - 2. Ford-4 Fine Coal Particle Size Distribution Fraction um Weight (g) Yield (%) Cum. Yield % >600 6.20 1.28 1.28 600 - 500 9.50 1.96 3.24 500 - 425 19.00 3.93 7.17 425 - 300 74.00 15.31 22.48 300-212 90.20 18.66 41.14 212-150 56.60 11.71 52.84 150-75 166.50 34.44 87.28 -75 61.50 12.72 100.00 Sum 483.50 100.00 Table Apdx - 3. Ford-13 Fine Coal Particle Size Distribution Fraction Weight Yield Cum. Yield um (g) (%) % >600 8.40 1.82 1.82 600 - 500 9.80 2.12 3.93 500 - 425 20.70 4.47 8.41 425 - 300 74.50 16.10 24.50 300-212 90.20 19.49 43.99 212-150 56.60 12.23 56.22 150-75 142.80 30.86 87.08 -75 59.80 12.92 100.00 Sum 462.80 100.00 Table Apdx - 4. Size Distribution of Ford-4 Ultrafine Coal by Wet Grinding Size Cum. Wt. Wt. (um) (%) (%) -3.00 1.00 1.00 3.00-6.00 6.56 5.56 6.00-9.00 17.72 11.16 9.00-12.00 35.06 17.34 12.00-15.00 53.77 18.71 15.00-18.00 68.25 14.48 18.00-21.00 78.22 9.97 21.00-24.00 85.70 7.48 " 24.00-27.00 90.23 4.53 27.00-30.00 93.46 3.23 30.00-36.00 97.00 3.54 36.00-45.00 99.09 2.09 +45.00 100.00 0.91 Total 100.00 229 Table Apdx - 5. The size distribution of Ford-4 coal sample used in the abstraction tests Fraction Weight Cum. Wt. (pm) (%) (%) 600 - 425 6.32 6.32 425 - 300 16.43 22.75 300-212 20.03 42.78 212-150 12.57 55.35 150- 106 19.59 74.94 106 - 75 17.38 92.32 75-45 7.68 100.00 Sum 100.00 Table Apdx - 6. The size distribution of silica sample used in the abstraction tests Fraction Weight Cum. Wt. (pm) (%) (%) 600 - 425 28.80 28.80 425 - 300 23.48 52.28 300-212 15.25 67.53 212-150 15.40 82.93 150-106 8.05 90.98 106 - 75 6.36 97.34 75-45 2.66 100.00 Sum 100.00 230 Figure Apdx - 2. Zeta-potentials of UBC-1 latex particles H20 pH AGENTS FLOCCULANTS DISTIL- H20 I SHMP HIGH SPEED BLENDER 20,000 RPM BEAKER WITH BAFFLES 900 ML 2% SOLID 1 STANDARD SIX-PADDLE MIXER CONDITIONING AT 350 RPM I STANDARD SIX-PADDLE MIXER SPEED CONTROLLABLE 350 RPM 3 MIN. 30 MIN. 2 MIN. STANDARD SIX-PADDLE MIXER SPEED ADJUSTED TO A CERTAIN VALUE t FILTRATION DEVICE 3 MIN. Figure Apdx - 3. Flocculation and filtration test procedures FILTER V A C U U M B A L A N C E 0 V E N B A L A N C E Figure Apdx - 4. Filtration and dewatering system set-up 233 Table Apdx - 7. The Size of Oil-droplets in Various Emulsions No. Oil Surfactant Oil-Droplet Type Amount added Cone, in aque.phase Initial addition phase in Size (u.m) Mean size d50 Elec. charge 0 Kerosene none none 0 Mechanically emulsified 2-100 20 -1 Kerosene DD Amine 5.4x10"4M Surfactant in water * 0.4-4.0 1.43 + 2 Kerosene DD Amine 1 % of oil wt. Surfactant in oil # 0.4-3.5 1.42 + 3 Kerosene DD Amine 5.4x10"5M Surfactant in water 0.5-5.8 1.95 + 4 Kerosene DD Amine 0.1% of oil wt Surfactant in oil 0.4 - 3.8 1.64 + 5 Kerosene C16-18 Amine 3.9x1 O^M Surfactant in water 0.4-3.8 1.40 + 6 Kerosene C16-18 Amine 1% of oil wt. Surfactant in oil 0.4-3.0 1.40 + 7 Kerosene Ci6-i8Amine 3.9x10"5M Surfactant in water 0.5-9.0 2.85 + 8 Kerosene Ci6-i8Amine 0.1% of oil wt Surfactant in oil 0.4-3.8 1.41 + 9 Kerosene DD Sulfate 3.4x1 O^M Surfactant in water 0.4 - 3.7 1.46 -10 Kerosene Cetyl Sulfate 2.7x1 O^M Surfactant in water 0.4 - 3.7 1.34 -11 Kerosene CO-610 U x l O ^ M Surfactant in water 0.4-11.0 4.02 -* Surfactant in water - In emulsification, the surfactant was added in water, and oil was then emulsified in the aqueous solution. # Surfactant in oil - In emulsification, the surfactant was first dissolved in oil, and the oil phase was then emulsified in the water. SIZE (um) Figure Apdx-5. Size distribution of kerosene droplets emulsified in a 5.4xl0~4 M dodecyl amine solution 0 0.5 1 1.5 2 2.5 3 3.5 4 PARTICLE SIZE (um) Figure Apdx-6. Size distribution of kerosene droplets obtained by emulsifying kerosene containing 1 wt% of dodecyl amine in aqueous solution PARTICLE SIZE (um) Figure Apdx-7. Size distribution of kerosene droplets emulsified in a 5.4xl0"5 M dodecyl amine solution PARTICLE SIZE (um) Figure Apdx-8. Size distribution of kerosene droplets obtained by emulsifying kerosene containing 0.1 wt% of dodecyl amine in aqueous solution Figure Apdx-9. Size distribution of kerosene droplets emulsified in a 3.9xlCH M C 1 6 . 1 8 amine solution SIZE (um) gure Apdx-10. Size distribution of kerosene droplets obtained by emulsifying kerosene containing 1 wt% of C 1 6 . 1 8 amine in aqueous solution Figure Apdx-11. Size distribution of kerosene droplets emulsified in a 3.9xl0"5 M C 1 6 . 1 8 amine solution SIZE (um) Figure Apdx-12. Size distribution of kerosene droplets obtained by emulsifying kerosene containing 0.1 wt% of C ] 6 . i 8 amine in aqueous solution 238 SIZE (um) Figure Apdx-14. Size distribution of kerosene droplets emulsified in a 2.7xl(H M Cetyl sulfate solution Figure Apdx-15. Size distribution o f kerosene droplets emulsified i n a 1 .7xl0" 4 M nonylphenoxy polyethanol (CO-610) solution Apdx-16. Calculation of Surface Aera of Ford-4 Ultrafine Coal Based on the Data Shown in Table Apdx-8: Sv = To ta l surface area / To ta l V o l u m e = (C2/C3)-(l/dvs) 1/dvs = Z (fi /di) Sw = T o t a l surface area / To ta l weight = Sv /p = (Ci/d)- (l/p)-2 (fi /di) where Sv - Speci f ic surface area per unit vo lume ; Sw - Spec i f ic surface area per unit weight ; C2 - Surface area shape factor; C 3 - V o l u m e shape factor; di - Average dimeter o f each fraction o f particles; f - W e i g h t ratio o f each fraction; p - s o l i d particle density. The specif ic surface area is calculated as fo l lows : 240 Table Apdx-8. Calculation of Surface Area of the Ford-4 Ultrafine Coal di W t % fi 1 /d i f i / d i 0.05 0.07 0.0007 20.00 0.014 0.15 0.14 0.0014 6.67 0.009 0.25 0.34 0.0034 4.00 0.014 0.35 0.41 0.0041 2.86 0.012 0.45 0.47 0.0047 2.22 0.010 0.55 0.61 0.0061 1.82 0.011 0.65 0.74 0.0074 1.54 0.011 0.75 0.95 0.0095 1.33 0.013 0.85 1.15 0.0115 1.18 0.014 0.95 1.22 0.0122 1.05 0.013 1.25 1.69 0.0169 0.80 0.014 1.75 2.16 0.0216 0.57 0.012 2.25 2.57 0.0257 0.44 0.011 2.75 3.04 0.0304 0.36 0.011 3.25 3.38 0.0338 0.31 0.010 3.75 3.92 0.0392 0.27 0.010 4.50 4.59 0.0459 0.22 0.010 5.50 5.07 0.0507 0.18 0.009 6.50 5.47 0.0547 0.15 0.008 7.50 5.74 0.0574 0.13 0.008 8.50 6.22 0.0622 0.12 0.007 10.0 6.62 0.0662 0.10 0.007 11.5 6.69 0.0669 0.09 0.006 13.0 6.76 0.0676 0.08 0.005 15.0 6.62 0.0662 0.07 0.004 18.0 6.42 0.0642 0.06 0.004 22.5 5.85 0.0585 0.04 0.003 27.0 5.41 0.0541 0.04 0.002 33.0 3.38 0.0338 0.03 0.001 40.0 1.35 0.0135 0.03 0.000 44.0 0.34 0.0034 0.02 0.000 45.0 0.61 0.0061 0.02 0.000 C2/C3 1/p Sw S U M 100.00 1.00 0.2644 0.833 (cmVg) 15417.1 241 Table Apdx-9. Electrode Potential Measurements of Dodecyl Amine Chloride Solutions at Different Concentrations using the Selective Electrode (Figure 5.1.3-1) at pH 5.92-5.93 Test 1 Test 2 Dodecyl amine chloride concentrations (mg/l) 0.01 0.1 1 10 100 0.01 0.1 1 10 100 Readings 1 94.3 90.7 81.3 34.9 -11.7 94.5 84.4 82.2 35.5 -11 .8 2 98.4 89.4 81.7 35.3 -12.0 91.5 85.4 81.3 34.9 -11 .7 3 mv 96.7 89.3 79.7 35.5 -11.8 94.4 83.4 81.4 35.6 -12 .1 4 97.6 87.8 81.1 34.7 -12.1 95.7 84.4 80.8 35.3 -11 .7 5 95.4 89.3 81.4 35.3 -11.9 93.4 85.4 81.1 34.9 -12 .0 Average Std Dev 96.48 1.477 89.30 0.919 81.04 0.697 35.14 0.294 -11.90 0.141 93.90 1.404 84.60 0.748 81.36 0.467 35.24 0.294 -11.86 0.162 Table Apdx-10. The Data for Calibration Curve of UBC-1 Latex Concentration against Solution Transmittance (Light Wavelength 800 nm) UBC-1 Cone. Transmittance Ln(transmittance) (ppm) Read#1 #2 #3 Average Read#1 #2 #3 Average Std Dev 0.2 96.0 95.9 96.0 95.97 4.564 4.563 4.564 4.564 0.0005 2 91.6 91.4 91.4 91.47 4.517 4.515 4.515 4.516 0.0010 8 72.0 72.3 72.1 72.13 4.277 4.281 4.278 4.279 0.0017 12 59.4 59.6 59.2 59.40 4.084 4.088 4.081 4.084 0.0027 20 43.6 43.6 43.4 43.53 3.775 3.775 3.770 3.774 0.0022 30 32.7 31.7 31.5 31.97 3.487 3.456 3.450 3.465 0.0163 40 18.9 19.0 19.2 19.03 2.939 2.944 2.955 2.946 0.0065 r squared 0.9936 0.9964 0.997 0.996 a constant 4.5866 4.5873 4.584 4.586 b coefficient -o:o4 -0.04 -0.04 -0.04 For Ln(T)= a+ b(C) 242 Table Apdx-11 The Data for Calibration Curve of Kero-DDA Emulsion Oil Concentration against Solution Transmittance (Light Wavelength 800 nm) Kero-DDA oil cone. Transmittance Ln(transmittance) (ppm) Read#1 #2 #3 Average Read#1 #2 #3 Average Std Dev 20 92.4 92.7 92.4 92.5 4.526 4.529 4.526 4.527 0.0015 50 81.5 81.9 82.0 81.8 4.401 4.405 4.407 4.404 0.0026 100 66.5 66.3 66.7 66.5 4.197 4.194 4.200 4.197 0.0025 200 43.1 43.8 43.6 43.5 3.764 3.780 3.775 3.773 0.0068 400 18.4 18.1 18.1 18.2 2.912 2.896 2.896 2.901 0.0077 600 8.8 8.8 8.8 8.8 2.175 2.175 2.175 2.175 0.0000 1000 1.6 1.8 1.7 1.7 0.470 0.588 0.531 0.531 0.0481 r squared 0.9997 0.9993 0.9995 0.9996 a constant 4.5621 4.5876 4.596 4.5951 b coeff -0.004 -0.004 -0.004 -0.004 For Ln(T)= a+ b(C) Table Apdx-12. Size Distribution Measurements for Kerosene Droplets Emulsified in a 2.7x10-4 M Cetvl Sulfate Solution Test 1 (Fresh) Test 2 (after 24 hrs) Particle Differential Cumulated Particle Differential Cumulated size (um) volume (%) volume (%) size (um) volume (%) volume (%) 0.447 2 0.01 0.47 2 0.14 0.603 14 2.12 0.667 22 4.07 0.775 40 9.07 0.857 52 14.02 0.947 62 20.16 1.046 72 27.39 1.216 88 40.43 1.279 93 45.28 1.413 99 55.59 1.486 99 60.92 1.562 99 66.25 1.727 89 76.43 1.815 81 81.04 1.908 72 85.14 2.109 49 91.61 2.218 40 94.03 2.709 9 99.11 2.848 7 99.49 3.659 0 100.00 3.725 0 100.00 Geo mean 1.303 1.308 Median 1.34 1.34 Arith mean 1.388 1.39 Std dev 0.496 0.497 243 Table Apdx-13. Measurements of Filtration Rates and Cake Moisture Contents of Ford-4 Ultrafine Coal with the Use of Different Additives Cake moisture content % Relative experiment deviation Additives Dosage Test #1 #2 #3 Average Test #1 #2 #3 0 38.25 38.94 38.97 38.72 -1.214 0.568 0.646 FR-7A 50 41.97 42.80 42.52 42.43 -1.084 0.872 0.212 (g/t) 100 43.94 44.52 45.88 44.78 -1.876 -0.581 2.456 300 43.38 43.50 43.26 43.38 0.000 0.277 -0.277 500 41.83 42.05 42.06 41.98 -0.357 0.167 0.191 0 38.25 38.94 38.97 38.72 -1.214 0.568 0.646 PAM 50 40.80 40.25 39.79 40.28 1.291 -0.074 -1.216 (g/t) 100 41.33 40.98 40.21 40.84 1.200 0.343 -1.543 300 40.88 41.26 41.79 41.31 -1.041 -0.121 1.162 500 41.40 42.13 41.03 41.52 -0.289 1.469 -1.180 0 38.88 40.05 40.32 39.75 -2.189 0.755 1.434 0.5 43.50 42.90 40.20 42.20 3.081 1.659 -4.739 Kero-DDA 1 48.55 48.46 47.68 48.23 0.663 0.477 -1.140 mulsion % 2 47.12 46.99 46.50 46.87 0.533 0.256 -0.789 3 47.88 48.25 45.62 47.25 1.333 2.116 -3.450 4 46.87 47.26 48.82 47.65 -1.637 -0.818 2.455 0 38.88 40.05 40.32 39.75 -2.189 0.755 1.434 100 39.55 39.47 39.72 39.58 -0.076 -0.278 0.354 DD Amine 200 41.75 39.10 38.76 39.87 4.715 -1.931 -2.784 (g/t) 400 41.78 39.66 39.01 40.15 4.060 -1.220 -2.839 1000 40.15 38.11 40.84 39.70 1.134 -4.005 2.872 2000 38.55 39.50 41.11 39.72 -2.946 -0.554 3.499 Filtration rate (ml/min) Relative experiment deviation Additives Dosage Test #1 #2 #3 Average Test #1 #2 #3 0 27.90 27.50 26.50 27.30 2.198 0.733 -2.930 FR-7A 50 76.60 80.10 78.80 78.50 -2.420 2.038 0.382 (g/t) 100 78.50 78.80 78.50 78.60 -0.127 0.254 -0.127 300 79.20 81.20 75.10 78.50 0.892 3.439 -4.331 500 64.50 65.30 69.40 66.40 -2.861 -1.657 4.518 0 27.90 27.50 26.50 27.30 2.198 0.733 -2.930 PAM 50 43.90 45.80 43.20 44.30 -0.903 3.386 -2.483 (g/t) 100 45.80 46.90 45.90 46.20 -0.866 1.515 -0.649 300 40.80 42.80 42.70 42.10 -3.088 1.663 1.425 500 39.20 38.40 40.90 39.50 -0.759 -2.785 3.544 0 27.90 27.50 26.50 27.30 2.198 0.733 -2.930 0.5 50.50 50.80 49.30 50.20 0.598 1.195 -1.793 Kero-DDA 1 61.50 62.80 61.70 62.00 -0.806 1.290 -0.484 mulsion % 2 73.20 74.50 76.70 74.80 -2.139 -0.401 2.540 3 86.50 87.80 88.20 87.50 -1.143 0.343 0.800 4 101.50 99.55 98.80 99.95 1.551 -0.400 -1.151 0 27.90 27.50 26.50 27.30 2.198 0.733 -2.930 100 39.10 36.40 37.90 37.80 3.439 -3.704 0.265 DD Amine 200 43.50 42.80 47.20 44.50 -2.247 -3.820 6.067 (g/t) 400 45.50 44.80 48.30 46.20 -1.515 -3.030 4.545 1000 51.20 52.30 47.10 50.20 1.992 4.183 -6.175 2000 52.10 55.20 53.20 53.50 -2.617 3.178 -0.561 CO • H O ID Q O CC CL < o UJ a co Q < CO r-< o cc m ID - J LL o g H O ZD • Q O CC CL M o o CO w o ° CD « _J X o o o o cc UJ r-<-o o < cc o 1 A 1 < x (-x a. < 2 LU z LU OO o cc LU O CD _i cc A O CD X cc 00 i X X 3 o n + o es O c j 1-O CD o o i ONlNHOdOAH 1 A o < cc DNLXVMHQ N0II0VmX3 1 1 o CD > 1-T CD 2 D D LU 2 3 X o CL l s i a OVA i s i a m v loa mv CO o .£2 CD ' o CO co .rz o TJ c to cu CD to o to O to to LU X TJ < CD Table Apdx-14. 245 MASS SPECTROMETER ANALYSIS - DECOMPOSITION-(//1156) OIL • OIL OIL OIL LIGHT 0 I L 805 846 2300 2600 . • FO 904 SATURATES 87.2 83.2. 78.1 69.5 16.2 PARAFFINS 26.6 26.4 15.5 • 9.0 10.1 1-RING 17.4 19.1 21.0 22.5 4.3 • 2-RING 15.1 15.6 16.3 15.0 1.5 3-RING 8.8 8.9 11.5 11.1 0.2 4-.RING ' y.4 7.8. 7.2 5.7 5-RING 5.2 3.2 . 3.4 2.2 6-RIN.G , 2.7 1.3 1.6 0.9 ' MONO Ar 0 0.9 1.7 3.1 AROMATICS 12.0 16.5 18.9 28.5 63.8 80 ALKYL BENZENES 3.5 4.8 3.8 5.2 6.9 20 NAPHTHENO-AROMATICS 4.3 6.3 6.2 9.4. • S.5 MONONAPHTHENE BENZENES 2.3 • 3.3 3.3 4.6 6.8 DINAPHTHENE BENZENES 2.0 2.9 2.9 4.8 2.7 TWO RING AROMATICS 1.5 2.6 3.0 5.2 61.8 28 NAPHTHALENES 0.5 0.9 0.9 1.5 50.6 ACENAFHTHENES 0.6 0.9 1.1 1.9 7.3 FLUORENES 0.5 0.8 1.0 1.8 4.0 THREE RING AROMATICS 0.3 - 0.5 1.0 1.8 2.4 io PHENANTHRENES 0.2 0.3 0.4 0.8 NAPHTHENEPHENANTHRENES 0.1 0.2 0.6 1.0 FOUR+ RING AROMATICS 0.6 0.4 0.6 0.7 PYRENES 0.2 0.1 0.2 0.3 CHRYSENES 0.2 0.1 0.2 0.1 PERYLENES 0.1 0.1 0.1 0.2 DIBENZANTHRACENES 0.1 0.1 0.1 0.1 BENZOTHIOPHENES 0.6 0.5 0.9 1.1 3.1 MULTIRING-S-AROMATICS 0.3 0.4 0.4 0.7 DIBENZOTHIOPHENES 0.3 0.3 . 0.3 0.5 NAPHTHOBENZOTHIOPHENES 0 0.1 0.1 0.2 UNIDENTIFIED AROMATICS 0.9 1.1 3.1 4.4 22 * Reference 2 Table Apdx-14 (continues) 246 PHYSICAL AND CHEMICAL INSPECTIONS OF QJL COMPONENTS ]' OIL 805 OIL 846 OIL 2300 OIL ' 2600 LIGHT FO HEAVY FO DENSITY, 15C, Kg/M3 861.2 873.5 881.1 897.9 932.1 .1123.5 K I N E M A T I C V I S C O S I T Y cSt@ 4 0 C 18.0 29.5 107 310 ' 3.41 1230 cSt @ 100C 3.75 5.0 11.2 22.3 S U S @ 1 OOF 95.3 151 560 1635. S U S @ 210F ' 38.9 •43.1 65.0 113 V ISCOSITY INDEX 92 91 90 88 C O L O R , A S T M <1.0 1.0 1.5 4.0 1.0 F L A S H POINT, C 195 200 238 .279 76 110 F L A S H POINT, F 383 392 460 534 169 230 P O U R POINT, C -15 -12 -3 3 ' -27 12 P O U R POINT; F . 5 10 27 37 ' -17 54 ANILINE POINT, C, 98 103 115 122 ANIL INE POINT, F 208 217 239 252 REFRACT IVE INDEX, 20C. •1.473 1.479 1.485 1.494 REFRACTIVITY INTERCEPT 1.045 1.044 1.046 1.043 VISCOSITY GRAVITY C O N S T A N T 0.814 0.524 ' 0.810 0.818 C L A Y G E L A N A L Y S I S S A T U R A T E S 80.8 83.5 82.1 70.4 20.3 A R O M A T I C S 18.9 16.3 17.3 27.4 75 .7 POLARS" 0.3 0.2 0.6 2.2 UV ABSORPTIVITY @ 260 nm • 0.43 0.85 0.97 2.4 C A R B O N T Y P E A N A L Y S I S A R O M A T I C C A R B O N , Ca 3.7 4.9 6.2 7.7 NAPHTHENIC C A R B O N , Cn 29.6 29.6 25.3 24.4' PARAFFINIC C A R B O N , Cp 66.7 65.5 68.5 • 67.9 M O L E C U L A R WEIGHT 350 395 520 600 S U P H U R , W T % 0.05 0.09 0.11 0.14 0.31 1.34 B A S I C NITROGEN, ppm 2 6 50 164 . 42 97 M E T A L S , ppm Cu 2 Fe 3 5 Va - 2 AJ 40 SI 35"' ASH, VYT% 0.1.5 DISTILLATION, C IBP 320 334 378 396 140 5 0 % off 395 412 494 544 287 F B P • 472 510 589 640 393 C A R B O N CONTENT, WT% 82.0 82.3 82.5 83.0 90.2 93 H Y D R O G E N CONTENT , WT% 18.0 17.6 17.4 16.8 10.1 6.9 

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