Particle fracture is the elementary process that governs comminution. In industrial machines particle breakage occurs mainly through three mechanisms: impact, abrasion and attrition. Of these mechanisms, impact breakage is known to be the most basic form of particle size reduction. Comminution devices are highly inefficient, as the energy used for particle breakage relative to that consumed by the equipment is low and reported to be between 1-2 %. As such, understanding the fundamentals of particle fracture is crucial for the development of energy efficient particle size reduction methods. Research done towards investigating particle fracture under impact loading has led to the development of several devices which include the twin pendulum device, drop weight tester, Split Hopkinson Pressure Bar, Rotary Breakage Tester and the Short Impact Load Cell. In this study the Short Impact Load Cell (SILC) was used to conduct bed breakage experiments on partially confined particles. Breakage tests using this device were conducted by vertically releasing a steel ball of known mass onto a bed of particles from a known height. The bed rested on a steel rod which was fitted with strain gauges to measure the particle response to impact loading. Tests were conducted on two ores, blue stone and UG2, to investigate the effect of three variables: steel ball mass, drop height and bed depth on the breakage behaviour of particles. The effect of each variable was investigated by evaluating the peak forces obtained, the particle fracture energy and the degree of particle breakage attained. For both ores it was found that the peak force increased linearly with increasing steel ball mass and drop height, and it was found that the drop height had a greater effect on the peak force than the steel ball mass. The maximum peak forces were obtained at one layer of particles and increasing the bed depth generally led to a reduction in the peak force. An exponential relationship was found between the peak force and bed depth, where the peak force decreased with increasing bed depth. It was found that the blue stone particles did not break at the range of input energies used in this work, therefore no fracture energy results were reported for blue stone. The fracture energy values for UG2 were low, where the maximum energy used for particle fracture was 2.7 % of the input energy. There was no direct correlation between the fracture energy and the steel ball mass, drop height and bed depth; however it was found that the bed depth had a larger effect on the fracture energy compared to both the steel ball mass and drop height. The greatest amount of energy used for fracture was generally obtained at the largest input energies using the 357 and 510 g balls. The optimum drop height which resulted in the highest fracture energy was generally found to be either 240 or 300 mm. A bed depth of five layers was found to be the optimum bed depth that allowed for the highest amount of energy to be utilized for breakage. No breakage results were obtained for blue stone due to the hardness and stiffness of the ore. For UG2, tests conducted at the same bed depth showed a trend in which the breakage initially increased greatly with increasing input energy; however at larger input energies the breakage obtained approached a constant value. Although the input energy was varied by changing both the steel ball mass and the drop height, the results showed that the degree of breakage was more dependent on the steel ball mass compared to the drop height. For all tests conducted, the maximum breakage was obtained at one layer of particles and increasing the bed depth led to a decrease in the breakage obtained. The results showed that the fracture energy and the degree of breakage were not directly related. It was found that there is an optimum amount of energy utilized for fracture that leads to the greatest breakage, where an in increase in the energy beyond the optimum point does not significantly affect the breakage obtained.

The Bushveld complex is the largest ore body in the world hosting platinum group elements (PGEs). It is a stratified orebody with three major reefs namely, the Merensky reef, UG2 reef and the Platreef. Platinum and palladium are the most abundant PGEs found in the Bushveld complex. They occur in the form of minerals/mineral phases with elements such as sulphur, tellurium, arsenic and iron. These minerals/mineral phases are associated with base metal sulphides occuring along grain boundaries. Unlike the Merensky and UG2 reef, the Platreef is almost barren of PGE sulphides and the distribution of base metals sulphides and their association with PGMs is erratic. Froth flotation targeted at the recovery of base metal sulphides is implemented in PGM concentrators to concentrate PGMs. Flotation of sulphide minerals is achieved with the use of thiol collectors to create hydrophobicity, and copper sulphate is often used to improve hydrophobicity and therefore recovery. Sodium ethyl xanthate (SEX) and sodium diethyl dithiophosphate (DTP) are commonly used as collectors on PGM concentrators. The erratic mineral variations in the Platreef ore, however, raise the question of the effectiveness of the application of sulphide mineral flotation techniques on this ore. Previous work by Shackleton, (2007) investigated the flotation of PGE tellurides, sulphides and arsenides. The study highlighted that the mechanisms with which these minerals interact with collectors and with copper sulphate was poorly understood. It is as a result of the findings of Shackleton's work that this study aims to elucidate the fundamental interactions of telluride and sulphide PGMs with thiol collectors and with copper sulphate. Subsequently this work also aims to compare the behaviour of these reagents on sulphide PGMs and telluride PGMs.

Due to the depletion of coarser grained ores, more mineralogically complex ores are being treated. These complex ores usually have finer grained valuable minerals. Liberation of these finer grained valuable minerals lies in grinding finer. Grinding to these fine sizes is energy intensive and using standard ball mills are energy inefficient at these sizes (P80 < 75μm). Therefore, stirred mills are becoming increasingly prevalent in the mineral processing industry. In order to optimize these mills, the effects and mechanisms of the significant variables need to be understood. This project investigated operating parameters against performance in a laboratory scale vertical stirred mill (Deswik mill), in an ultrafine grinding (UFG) application of MG2 reef in the bushveld igneous complex. The operating variables that were investigated are stirrer speed, solids concentration, media size and media filling. The Kwade stress energy model was tested on the grinding results. The grinding performance was quantified in two ways, i.e. grinding efficiency and grinding rate. The grinding performance for this study was also investigated through a statistical analysis. The experiments was designed using a face centred central composite design (FCCD) and the results was statistically analysed using a design of experiments (DOE) software.

Froth flotation is a separation method used for the beneficiation of a considerable portion of the world's mineral ores. The majority of flotation occurs in mechanical flotation cells, where effective gas dispersion is a primary requirement for particle-bubble contacting. Due to the mineralogical complexity of ores, it is required that particles be ground even finer to liberate valuable minerals. Mining operations tend to run flotation circuits at fairly high solids concentrations in order to maximise residence time, accommodate higher tonnages and limit water consumption. Mineral slurries processed at fine particle sizes and high solids concentrations have been shown to exhibit non-Newtonian rheological behaviour. The effect of slurry rheology on gas dispersion in a 100 litre mechanical flotation cell was investigated by varying the solids concentration. The study was conducted using kaolin, Bindura nickel and Platreef slurries. All three ores displayed typical non- Newtonian rheological behaviour where the slurry yield stress and viscosity increased exponentially with solids concentration. Bubble size varied from 0.55 to 1.10 mm for all the ores tested. At low solids concentration bubble size was found to decrease with impeller speed, a characteristic trend that was expected. At moderate solids concentrations bubble size was found to either increase/remain relatively constant with impeller speed; this trend was also expected. Unexpectedly, at the highest solids concentration, a dramatic decrease in bubble size was observed. This unexpected drop in bubble size was attributed to slurry rheology. It was also observed that there was a slight increase in bubble size at the highest solids concentration with increasing impeller speed. This increase was attributed to a trade-off relationship between the rheology of the slurries and the existing hydrodynamics (as a result of the rotating impeller). Gas hold-up varied from 2 to 15% across all the ores tested. At low solids concentrations gas hold-up increased with impeller speed as expected. At moderate solids concentration gas hold-up was viewed to either increase/remain relatively constant with impeller speed. A significant drop in gas hold-up was observed at the highest solids concentration. The gas hold-up however still increased with impeller speed albeit at a lower rate at the highest solids concentrations. This drop in gas hold- up at the highest solids concentration (along with the decrease in bubble size) was attributed to the effect of slurry rheology. At high solids concentrations, all three slurries (kaolin, Bindura nickel and Platreef) exhibit non-Newtonian behaviour illustrated by means of high viscosities and yield stresses. High viscosities result in turbulence damping in the cell which inhibits bubble break-up, resulting in larger bubbles and correspondingly lower gas hold-up. It was concluded in this study that the yield stress is the dominant rheological property due to the significant changes observed with increasing solids concentration. High yield stresses resulted in the formation of a 'cavern' of slurry around the impeller region. Within this 'cavern', high power intensities exist around the impeller where small bubbles are formed. However due to the formation of the 'cavern', the slurry in the bulk cell remains relatively stagnant. As a result small bubbles formed around the impeller remain localised in the 'cavern' and cannot be dispersed throughout the cell. This localization and poor dispersion of bubbles resulted in low gas hold-ups.

As the excessive extraction of high grade iron ore reserves has led to the rapid depletion of these ore bodies, there is a growing need to extract and upgrade low grade iron ores into more economically viable products with an iron content in excess of 50%. The beneficiation of low grade iron ores through the reverse cationic flotation procedure is gradually gaining popularity as a possible processing route of the future for South Africa's iron industry. Reverse cationic flotation employs a reagent suite consisting of an amine compound which functions as a quartz collector in addition to providing the frothing effect in the flotation system, and hydrolysed starch which serves to depress hematite during flotation. The aim of this project was to investigate the effect of using five amine collectors with different molecular structures on the flotation recovery of quartz and the entrainment of hematite in the flotation of a South African iron ore and a Brazilian iron ore. Laboratory batch flotation tests were conducted on both ore samples and the grade and recovery of hematite were recorded. The collectors were characterised through surface tension measurements and pKa value analysis. An attempt at using different polysaccharides as hematite depressants entailed the use of a CMC and a guar gum in batch flotation tests of the Brazilian iron ore.

Summary in English. / Bibliography: pages 96-100. / The extraction of copper from chloride solutions with some aliphatic oximes as well as some pyridine carboxylates was studied and the complexes formed in the organic phase identified. The association of octanal oxime in toluene was taken into account in the extraction studies and formation constants for dimers and trimers obtained. Copper (II), Nickel (II) and Cobalt (II) are extracted by a solvating mechanism in which only the neutral MCl2 species is extractable. This reaction is largely independent of pH. Studies were also carried out on the use of the commercial reagent ACORGA CLX-20, for the selective extraction of copper from a synthetic solution simulating leach liquors obtained by ferric chloride leaching of complex sulphide ores. Results show that separation of copper from iron is dependent on the amount of acid and chloride present in the aqueous phase.

Classification is an integral part of comminution operations that controls the performance of the circuit. Hydrocyclones are normally used to perform the classification function. They offer numerous advantages that include, the ability to handle high throughputs, low floor space occupation and relatively low capital and running costs. Despite these advantages, hydrocyclones are inherently inefficient classifiers as they are predominantly dependent on hydrodynamics to effect separation. This effect is more prominent in operations handling complex ores such as a dual-density ore, where the heavy fine particles are misplaced to the underflow and the lighter middling particles report to the overflow. Several attempts have been made to improve the separation efficiency of cyclones either by modification of the cyclone or use of multi-stage cycloning. Most of the results obtained from experimental and simulation studies have shown considerable improvements. Even though some have not yet found wide application in the minerals industry due to practical limitations related to control and unstable operations. More recently, fine screening has gained recognition in the classification role. This development has allowed the use of fine screens in closed-circuit grinding operations resulting in significant metallurgical and economic benefits. Screens provide a sharper cut at the desired size and reduce the fraction of fines bypassing classification compared to hydrocyclones but have capacity limitations at smaller apertures. In an effort to mitigate the classification challenges of both the hydrocyclone and fine screen, this study investigated the potential of combining the high throughput performance of the hydrocyclone operation and the high precision classification characteristics of fine screening to result in a hybrid classification circuit Plant scale tests were conducted using five different classification circuit configurations at an operational Base Metal Concentrator treating a polymetallic ore. The classification circuit configurations considered included (i) a two-stage hydrocyclone with primary underflow reclassification (ii) an inclined hydrocyclone, (iii) a fine screen and (iv) selected permutations of hybrid circuit designs that included a hydrocyclone-fine screen (2 stage) and two hydrocyclones-fine screen (3 stage) variants of the hybridised configurations. The efficiency curves and their respective key performance indicators were used to assess the performance of the circuit configurations tested. The results showed that classification circuits that included fine screens exhibited higher sharpness of separation compared to circuit configurations comprised of hydrocyclones. The fine screen configuration showed the sharpest separation while the hydrocyclone-fine screen hybrid configurations gave relatively higher separation efficiencies than the configurations with hydrocyclones only. The overall sharpness of separation values obtained for the two stage and three-stage hybrid circuits were 3.0 and 2.4, respectively. The two-stage hydrocyclone and inclined hydrocyclone circuits had sharpness of separation values of 1.7 and 0.5, respectively. The inclined hydrocyclone circuit configuration performed the poorest. Furthermore, the two-stage hybrid circuit showed a higher degree of separation compared to the three-stage hybrid configuration. However, it was observed that a finer corrected cut size was realised for the three-stage hybrid circuit design. The fishhook effect was seen at particle sizes less than 38μm for the configurations incorporating a fine screen and an inclined hydrocyclone. Notably, the effect appeared to be more pronounced in configurations involving a fine screen stage. The results have shown that application of hybrid classification configurations can improve the performance of classification circuits. In addition, reclassification of hydrocyclone underflow on fine screens will results in a sharper classification while reclassifying the overflow stream on fine screens will provide a clean circuit final product. An evaluation of the capital and operating costs associated with fine screens should be done to determine the economic feasibility of incorporating the units in conventional milling circuits.

There is a growing importance in the mineral processing industry to find ways which are economic and effective in improving the recovery of minerals in the flotation process. The focus of this study was on the recovery by flotation of minerals found in the Merensky reef, which is one of the major reefs in the Bushveld complex. In that reef, base metal sulphide (BMS) minerals are commonly associated with PGMs and this has an effect on the way in which these minerals are concentrated by flotation (Vermaak et al. 2004; Wiese et al. 2005b; Miller et al. 2005; Schouwstra et al. 2000).A major problem in this process has been reported to be losses of valuable minerals (PGMs) associated with the loss of BMS (Wiese et al. 2005b), during flotation. The present investigation has focused on studying the relationship between the flotation of sulphide minerals using xanthates as collectors and the electrochemical properties of the flotation system. It is well known that electrochemical mechanisms in flotation systems have a major influence on flotation since the reactions occurring at the mineral/solution interface are of critical importance in the process (Woods, 1971).The aim of this study was to investigate the extent to which there was a relationship between the electrochemical reactions occurring in this ore which could indicate the effectiveness of the flotation process. The electrochemical reactions were studied by determining the redox potential changes occurring when various changes were made. These were the length of the alkyl chain length of the xanthate collector, changing the pH or using various chemical reagents to change the potential of the system. It was found from the rest potential measurements, that collectors of different chain length have different extents of interaction with mineral surface. A greater interaction, which is indicated by a greater change in the mixed potential after addition of the collector, is considered to be indicative of a greater adsorption of the collector at the mineral surface. It was hypothesized that this stronger adsorption by collectors of longer alkyl chain length would result in improved flotation performance. However, this was not observed to be the case and that was consistent with previous results on the relationship between the recovery of sulphide minerals in the Merensky ore and xanthates of different chain lengths. Thus it was shown that there was no correlation between the interactions between collectors of different alkyl chain lengths as determined through electrochemical studies and the flotation performance of valuable minerals under the tests conditions used.

Due to continually declining ore grades, increasing mineralogical complexity, and increasing metal demand, models for the design and optimisation of minerals processing operations are of critical importance. These models do not currently incorporate particle shape, which, although rarely quantified, is known to affect numerous unit operations. Automated Scanning Electron Microscopy (Auto-SEM-EDS) is a widely used tool for mineralogical analysis. It also provides an opportunity for simple, quantitative and mineral-specific shape characterisation. Existing mineralogical databases could therefore become useful resources to facilitate the incorporation of shape effects in minerals processing models. A robust Auto-SEM-EDS shape characterisation methodology is required to ensure that the particle shape information in these databases is interpreted appropriately. For this work, a novel methodology for Auto-SEM-EDS shape characterisation was developed that is suitable for the analysis of fine particles (<75 μm). This involved testing the response of various shape descriptors to image resolution, and measurement with different devices and image processing routines. The most widely used shape descriptor in minerals processing, circularity, was found to be highly dependent on both image resolution and image processing settings, making it a poor choice for shape characterisation of fine particles. Roundness and aspect ratio were found to be more robust descriptors. However, in the interest of being able to compare particulate shape measurements across different studies, the precise definition of aspect ratio is important as variation in 'length' and 'width' definitions can significantly impact aspect ratio measurements. The possibility that preferential orientation of particles would introduce bias to the 2-D cross-sectional measurements was also addressed through comparison of roundness distributions measured from orthogonal cross-sections of a particulate sample mounted within a block of resin. The excellent repeatability of these measurements indicated that the particles were randomly orientated, and thus it can be inferred that 2-D measurements of a sufficient number of particles will be directly related to the particulate sample's 3-D properties. Roundness and aspect ratio were then used in conjunction to produce surface frequency distributions that allow for distinction between non-rounded particles that were smooth and elongated and non-rounded particles that were neither elongated nor smooth. Three applications of the shape characterisation methodology developed were then demonstrated, which highlighted some of the potential contributions that this methodology can make towards minerals processing. The applications were all based on a case study of the Upper Group 2 (UG2) Chromitite, a platinum group mineral (PGM) ore of key economic significance to South Africa.

Chalcopyrite is not only the most abundant of the copper sulphides, but also the most stable, making it recalcitrant to hydrometallurgical treatment processes especially in atmospheric leaching. Hence, pyrometallurgical processes are traditionally used to treat chalcopyrite concentrates. However, ore grades are falling and concentration processes are becoming increasingly costly, prompting need to revisit hydrometallurgical treatment processes (especially heap leaching), which are otherwise regarded as relatively economic and environmentally friendly. Key hydrometallurgical processes for chalcopyrite treatment are ferric sulphate, chloride and ammoniacal systems. The ferric sulphate system does not work well under atmospheric conditions, except in combination with thermophilic microorganisms, whereas the chloride system has only recently been evaluated more seriously for heap leach processes. The ammonia system remains relatively unexplored and most studies date back more than 40 years, but the system has considerable potential for further development. Ammonia systems can be effectively used to leach copper from chalcopyrite in the presence of an oxidant. The ammoniacal leaching system is heavily reliant on a good surface mass transfer system, hence it being widely studied in high pressure systems where oxygen was accepted to be the oxidant. Leach reactors were designed to use agitation systems which promote the abrasion of an iron based deposit layer thought to passivate the mineral surface. Most research on the ammonia leaching systems has previously been carried out in controlled or bulk leaching studies and only a few used electrochemical studies. A disconnect exits between the two approaches, resulting in different proposed fundamental reaction mechanisms and kinetic understanding. A fundamental electrochemical and controlled leach study of the oxidative leaching of chalcopyrite in ammoniacal solutions has been undertaken. The study covered the following aspects: a description of the mixed potentials, chemistry and kinetics of the anodic reaction, the cathodic reduction of the oxidants, the formation and effect of surface deposits and lastly a look at how results from electrochemical studies compare to those from the leaching of a similar mineral sample under similar solution conditions. A detailed study of the mixed potentials on a more or less pure chalcopyrite electrode has shown the redox reactions on the surface of the mineral to be controlled by the oxidation of chalcopyrite and reduction of copper(II). The presence of oxygen has been found to have no significant effect on mixed potentials in ammoniacal solutions in the presence of initial copper(II). Constant potential and potentiodynamic studies on the anodic reaction have shown the rate of the anodic reaction to increase with an increase in potential in a standard 1M ammonia/ammonium sulphate solution (which buffers at pH 9.6) in exponential fashion supporting conventional Butler-Volmer behaviour with a anodic transfer coefficient of 0.42 and a rate constant k* CuFeS2 of 0.0431 cms⁻¹. Increasing total ammonia increased the rate of reaction only at low concentrations; at higher concentrations increasing total ammonia had no effect on the anodic reaction. An increase of pH at fixed total ammonia concentration showed an increase in reaction rate, but the effect cannot clearly be discerned from the concomitant shift in relative proportion of free NH₃ and NH₄⁺. Coulometric studies have shown the oxidation reaction to proceed via the formation of a thiosulphate intermediate and this to be a 7-8 electron transfer reaction. A surface deposit layer consisting of iron, oxygen and small quantities of sulphur was formed and the sulphur component of this product layer was seen to be gradually depleted during leaching. Anodic currents were found to gradually decrease with time and this was linked to the growth of the surface deposit layer. However, the surface deposit layer did not passivate the anodic reaction; instead, it was proposed that the surface deposit layer adsorbed copper ions and displayed "ohmic" behaviour. The formation of the surface deposit layer was found to apparently promote the cathodic reduction of copper(II). While reduction of copper(II) was shown to be the primary reduction reaction, the presence of oxygen was seen to promote this reduction reaction through the regeneration of copper(II) in experiments that ran for longer time periods. An apparent accumulation of copper(I) on the mineral surface was seen to adversely affect the rate of the cathodic reaction and thus the overall rate of dissolution. The nature and morphology of the surface layer was found to be significantly influenced by the choice of cation in solution, which was thought to influence primarily the complexation/precipitation of ferric species forming near the surface. The degree of agitation during leach studies influences the rate of leaching due to the fragmentation of surface deposits, which are seen to slow the anodic reaction. A kinetic model has been developed for the anodic and cathodic reactions. This thesis presents significant new findings regarding the role of the copper(I)/copper(II) redox couple on the oxidative leaching of chalcopyrite. It also highlights the potentially limiting role of the cathodic reactions which have frequently been overshadowed by the focus on chalcopyrite oxidation reactions. Furthermore, the growth of a surface inhibiting layer which cannot be removed in heap leach systems due to the lack of mechanical agitation can now potentially be addressed by looking into the complexation and precipitation characteristics of cations in solution for ammoniacal leach systems.