A study has been made of the oxidation of molybdenite in solutions of hypochlorite. The reaction was found to be first order in reagent concentration with a reaction rate constant of 1.90 x 10⁻² min⁻¹ cm⁻² at pH 9 and 45°C. The oxidation rate increases with increasing pH and later decreases if the pH is raised above 9.5. An activation energy of 6.3 ± 0.8 kcal/mole was observed for this reaction which was controlled by either chemical or mixed diffusion and chemical reactions. Other factors studied were the agitation, surface area, surface characteristics and effect of sulfates, chlorides, chlorites and chlorates. It was observed that hypochlorite is capable of selectively leaching molybdenite from copper sulfides and positive results were obtained when applying this property to the extraction of molybdenite from copper rougher concentrates. The formation of oxides during drying and/or leaching of copper sulfides was found to be detrimental to the process because they act as catalysts in the decomposition reaction of hypochlorite. Rates of normal decomposition of the reagent were studied over the range 35 to 60°C and pH 7 to 10 and found to be negligible as compared to the rates of oxidation of molybdenite. The leaching of rougher concentrates with hypochlorite was found not to affect considerably the subsequent flotation operation. Data of on-site hypochlorite generator plants and factors that affect production and decomposition, have been presented with the purpose of proposing a process using the information obtained in this work for the extraction of molybdenite in the Copper Concentration Plants. / Applied Science, Faculty of / Materials Engineering, Department of / Graduate

Column leaching experiments were conducted to simulate in-situ leaching processes and to provide information on the mechanisms and extent of copper recovery, acid consumption, and chemical composition of recycled raffinate. Two 1.52-meter by 0.30-meter columns, each loaded with approximately 150 kilograms of copper oxide ore from the Florence porphyry copper deposit, were leached with a dilute sulfuric acid solution for 84 days. Computer simulation of the saturation state and aqueous mass transfer of predominant elements during the column tests was performed using EQ3NR, an aqueous solution speciation-solubility modeling code and the geochemical mass balance modeling code NETPATH. These modeling codes were used to quantify the amount of minerals dissolved and precipitated in the columns during the column leaching experiments. NETPATH mass balance models containing Cu-montmorillonite and chrysocolla as sources of copper match observations of pre-leach material. Cu-montmorillonite contributed 55% and 34% of the dissolved copper in the columns effluent.

A detailed investigation of the kinetics of dissolution of kaolinite clay in hydrochloric acid solution has been carried out. The rate of dissolution increases with calcination temperature, calcination time, and leach temperature. For example, almost 98% of the aluminum in the clay sample roasted at 540°C for 1 h can be dissolved in 20 min. The presence of fluoride ions in the leach solution significantly enhances aluminum dissolution. Changes in the pore structure of kaolinite clay due to calcination and hydrochloric acid leaching have been studied by a gas adsorption technique. The results obtained show that the surface area, macropore volume, and micropore volume of the leached residue increase with the extent of leaching up to a leach time of 1 h and thereafter decreases. Calcination at 500-750°C appears to destroy the structure of the clay but has no significant effect on the porous nature of the clay.

Thesis (MEng)--Stellenbosch University, 2015. / ENGLISH ABSTRACT: There is a lack of clear understanding of the rate of selenium (Se), arsenic (As) and sulphur (S) dissolution during caustic (NaOH) batch leaching of PGM-rich leach residue in the presence of oxygen. This has been a limitation in the optimisation of hydrometallurgical processes for the upgrading of PGM concentrates before refining the precious metals. Conditions to improve the rate of leaching of amphoteric elements while minimizing PGM losses were examined to enhance the performance of the leaching process. Development of intrinsic leaching rate equations represent the core of the overall batch leaching model developed in this study. The robustness of the model was assessed by its ability to accurately simulate the effects of changing operating parameters on the reaction extents. The effects of the interfacial oxygen mass transfer rate and temperature on the leaching rates were therefore also included in the overall model. The first part of the experimental program focussed on the interfacial oxygen mass transfer rate in the test autoclave. This enabled an accurate mathematical description of the interfacial mass transfer rate of the primary oxidant, diatomic oxygen (O2) molecule from the gas to the liquid phase. Mass transfer tests were conducted using the sodium sulphite method at 60°C, 100 kPa oxygen partial pressure and agitation speed of between 500 to 1000 rev/min. Cobalt(II) was used as the catalyst with a concentration range of 1 to 5 mg/L. Oxidation of amphoteric elements was investigated by leaching of PGM-rich leach residue (residue from sulphuric acid leaching of converter matte) in caustic solution. The test work was conducted to determine the intrinsic leaching rates in 0.125, 0.25 and 0.5 mol/L NaOH solutions in the 160° to 190°C temperatures range over a period of 6 hours. Oxygen partial pressure was maintained at 11 atm in the factorial experiments. The effect of oxygen partial pressure was quantified by conducting tests with oxygen partial pressure ranging from 7 to 16 atm. The intrinsic rate constants and activation energies derived from this test work were incorporated in the overall kinetic model to simulate the batch leaching profiles under real plant conditions. During the caustic pressure oxidation of amphoteric elements, the rate of oxidation was rapid during the first 10 minutes and decreased steadily over the course of experiment. The experimental results suggest that the oxidation kinetics are controlled by product layer diffusion with sulphur, selenium and arsenic (Arrhenius) activation energies of 31.8 kJ/mol, 26.1 kJ/mol and 10.7 kJ/mol respectively over the temperature range of 160 to 190°C. The reaction mechanism, as well as the observed kinetic behaviour, is most likely due to the base metal/PGMs hydroxide layer that formed as a result of precipitation. An increase in temperature increased the sulphur and arsenic reaction rates. The selenium reaction rate also increased as the temperature was increased from 160 to 175°C. A further increase in temperature above 175°C did not yield a significant increase in the reaction rate. An increase in the caustic concentration increased the reaction rates of all the elements. Increased oxygen partial pressure also improved the reaction rates, with the most significant change observed for sulphur oxidation; the extent of sulphur oxidation increased from 75 to 85% when oxygen partial pressure was increased from 7 to 16 atm. Reaction orders of 0.25, 0.12 and 0.21 with respect to hydroxide concentration and 0.37, 0.29 and 0.36 with respect to dissolved oxygen concentration were obtained for sulphur, selenium and arsenic respectively. Kinetic models were developed for sulphur, selenium and arsenic extraction. The sulphur and selenium simulation gave better agreement between the experimental and model predicted values, while the arsenic simulation gave a relatively poor prediction of the extractions. The caustic concentration had a notable effect on the dissolution of the PGMs. An increase in the caustic concentration increased the dissolution of platinum, palladium and ruthenium. Ruthenium dissolution also increased with an increase in temperature. To the contrary, platinum and palladium dissolution decreased with an increase in temperature. Rhodium and iridium precipitated and did not report in the solution phase while osmium could not be traced. The oxygen partial pressure did not have a significant effect on the dissolution rate of platinum, palladium and ruthenium. / AFRIKAANSE OPSOMMING: Daar is ‘n tekort aan die begrip van die tempo van seleen (Se), arseen (As) en swawel (S) oplosbaarheid gedurende bytsoda (NaOH) enkelladingsloging van platinum groep metaal (PGM)-ryk oorskot materiaal in die teenwoordigheid van suurstof. Hierdie inligting word benodig wanneer die optimisering van tipiese hidrometallurgiese prosesse wat PGM oorskot materiaal opgradeer verlang word. Hierdie bytsoda druklogingsproses vind tipies voor raffinering van die PGM metale plaas. Kondisies wat die tempo van amfoteriese element-loging verbeter, terwyl die PGM verliese geminimaliseer word, was in hierdie werk geondersoek om sodoende die effektiwiteit van die logingsproses te verbeter. Die ontwikkeling van intrinsieke logingtempo vergelykings vorm die kern van die algemene enkelladingsloging model wat ontwikkel was. Die robuustheid van hierdie model word geevalueer op sy vermoë om akkuraat die effekte van veranderende bedryfstelsel parameters op die logingstempo van betrokke reaksies te simuleer. Die effekte van suurstof tussenvlak massaoordrag en temperatuur was ook in die algehele model ingesluit. Die eerste deel van die eksperimentele program het gefokus op die suurstof tussenvlak massaoordrag in die outoklaaf. ‘n Akkurate wiskundige model wat die massaoordrag van die primêre oksidant, diatomiese suurstof (O2), van die gas fase na die vloeistof fase beskryf, was gebruik om die suurstof oordragtempo te kwantifiseer. Suurstof massaoordrag toetse het van die natrium sulfiet metode gebruik gemaak by 60°C, 100 kPa suurstof parsiële druk en tussen 500 en 1000 rev/min roerspoed. Kobalt(II) het gedien as katalis wat tussen 1 tot 5 mg/L gevarieer was. Die amfoteriese element oksidasie was volgende ondersoek deur die PGM-ryk oorskot materiaal te loog met bytsoda (wat stroomop onderwerp was aan swawelsuur loging van omskakelaar mat). Die toetswerk wou die intrinsieke logingtempo’s met 0.125, 0.25 en 0.5 mol/L NaOH oplossings by temperature 160 en 190°C oor 6 uur residensie tyd vasstel. Die suurstof parsiële druk was konstant gehou op 11 atm in hierdie faktoriale eksperimente. Die effek van suurstof parsiële druk was apart vasgestel, deur die suurstof parsiële druk te varieër vanaf 6 tot 16 atm. Die intrinsieke tempokonstantes en aktiveringsenergieë wat in hierdie toetswerk afgelei is, was in ‘n algehele kinetiese model ingekorporeer wat die enkellading logingsprofiele gesimuleer het by aanleg kondisies. Die tempo van oksidasie was vinnig in die eerste 10 minute en het geleidelik afgeplat, gedurende die bytsoda druk oksidasie van amfoteriese elemente. Die eksperimentele resultate suggereer dat produklaagdiffusie die oksidasie kinetika beheer met swawel, seleen en arseen (Arrhenius) aktiveringsenergieë as volg bereken in die temperatuur interval 160 tot 190°C: 31.8 kJ/mol, 26.1 kJ/mol en 10.7 kJ/mol. Die reaksie meganisme en kinetiese gedrag word hoogs waarskynlik veroorsaak deur die onedelmetaal/PGM hidroksied laag wat deur middel van presipitasie gevorm het. Temperatuur toename het die swawel en arseen se reaksietempo’s verhoog. Met seleen het die reaksietempo met temperatuur toename tussen 160 en 175°C ook verhoog, maar afplatting het by 175°C opwaarts plaasgevind. Oor die algemeen het die bytsoda konsentrasie die amfoteriese elemente se reaksietempo’s verhoog. Die verhoging van die suurstof parsiële druk het ook die reaksietempo’s verhoog. Swawel oksidasie het van 75 tot 85% verhoog vanaf 6 tot 16 atm, wat die mees noemenswaardige verandering was. Swawel, seleen en arseen reaksieordes van 0.25, 0.12 en 0.21 met hidroksied konsentrasie en 0.37, 0.29 en 0.36 met opgeloste suurstof konsentrasie het die beste paslyn op die eksperimentele data tot gevolg gehad. Hierdie data was gebruik om die kinetiese modelle van swawel, seleen en arseen ekstraksie te ontwikkel. Terwyl swawel en seleen ‘n goeie paslyn vir die eksperimentele data tot gevolg gehad het, kon passing van arseen ekstraksie nie ‘n goeie model oplewer nie. Varierende bytsoda konsentrasie het ‘n noemenswaardige effek op die PGM ontbinding gehad. Wanneer die bytsoda se konsentrasie vermeerder word, los daar meer platinum, palladium en rutenium op. Rutenium ontbinding het tydens ‘n temperatuur toename verhoog. In kontras het platinum en palladium ontbinding velaag tydens temperatuur toename. Rodium en iridium het gepresipiteer en was nie ontbind nie. Osmium kon nie gemeet word nie. Die suurstof parsiële druk het nie ‘n noemenswaardige effek op platinum, palladium en rutenium ontbinding gehad nie.

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A detailed study of the physical and chemical characteristics of a flue dust sample from a coal fired reverberatory furnace was carried out. The characterization studies revealed that cuprous ferrite, delafossite, is one of the major copper bearing constituents in flue dusts collected from copper reverberatory furnaces. A knowledge of the leaching characteristics of delafossite is thus essential to the development of hydrometallurgical techniques for copper recovery from flue dusts. The recovery of copper from synthetic delafossite was investigated using an ammoniacal carbonate solution. The results obtained indicate that the dissolution of delafossite in ammoniacal solution is very slow, but can be dramatically improved by a reductive roast prior to leaching. The kinetics of dissolution of delafossite has been found to strongly depend on the extent of reduction, particle size and stirring speed. It has also been found that applied oxygen pressure increases the rate of dissolution and accentuates the particle size effect. The experimental data seem to fit a film diffusion model. Limited studies have been carried out on the recovery of copper from a flue dust collected from a coal fired reverberatory furnace. Results from these studies have been compared with those obtained for synthetic delafossite.

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This study was aimed at recovering lithium from china clay waste using a combination of froth flotation, magnetic separation, roasting and leaching. The china clay waste produced by Goonvean Ltd contains about 0.84% Li2O and 0.36% Rb2O, present in some of the mica minerals. Among the mica minerals, zinnwaldite is the major source of lithium with smaller amounts being contributed by muscovite. The results of the flotation tests showed that the dodecylamine collector dosage had a greater effect on the recovery and grade of mica minerals to concentrate than pH over the range tested. It was found that a mica concentrate containing 1.45% Li2O, 0.55% Rb2O and 4.47% Fe2O3 could be produced at a recovery of 98.6%, 85.2% and 92.8% respectively. Mineralogical analysis of the flotation products showed that the concentrate consisted mainly of muscovite, zinnwaldite and kaolinite with minor amounts of K-feldspar and quartz. The tailing consisted of mainly quartz, K-feldspar and kaolinite with minor amounts of apatite, topaz, zinnwaldite and muscovite. Further upgrading of the concentrate was found to be possible using a wet high intensity magnetic separator producing a magnetic fraction containing 2.07% Li2O, 0.74% Rb2O and 7.42% Fe2O3 with a recovery of 73%, 67% and 77% respectively. A mineralogical analysis of the separation products showed that the magnetic fraction consisted of predominantly zinnwaldite with muscovite as the main contaminant. The non-magnetic fraction consisted of muscovite and kaolinite as the main minerals while zinnwaldite, K-feldspar and quartz were subordinate. Electron-microprobe analysis on individual mica grains have shown that zinnwaldite and muscovite contain on average a calculated Li2O content of 3.88% and 0.13% respectively. Lithium extraction from the concentrate is only possible after the lithium has been converted into a water-soluble compound. Thus, in order to convert the lithium in concentrate into a water-soluble compound, the gypsum and limestone lithium extraction methods together with the new method of using sodium sulphate were investigated. The process involved roasting a predetermined amount of lithium-mica concentrate with either gypsum, limestone or sodium sulphate at various temperatures and subsequently leaching the pulverised materials in water at 85oC. A lithium extraction efficiency of about 84% was obtained using gypsum at 1050oC while rubidium extraction was very low at 14%. It was found possible to extract about 97% Li and 16% Rb if the concentrate was roasted with sodium sulphate at 850oC. Processing the concentrate with limestone resulted in very low lithium extraction. Iron co-extraction was low in all cases. The XRD analysis of the gypsum and sodium sulphate roast-products showed that the water soluble lithium species were KLiSO4 and Li2KNa(SO4)2 respectively. Preliminary tests on the leach solution obtained by using sodium sulphate as an additive have shown that a Li2O3 product with a purity of > 90% could be produced by precipitation with sodium carbonate although more work is required to reach the industrial target of > 99%. The lithium carbonate obtained with Li2CO3 content of about 90% is still suitable for use in the glass and ceramic industries, and as feedstock for the production of high-purity lithium compounds. An economic evaluation of the proposed lithium carbonate production plant has indicated an annual rate of return on the investment before tax of 7.2%.