3. Electronic Theses and Dissertations (ETDs) - All submissions
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Item Improving the separation efficiency of hematite from slimes through selective flocculation(2019) Da Corte, CarlaThe prevalence and treatment of low grade, finely disseminated iron ore has resulted in the production of primary and secondary slimes that constitute potential resources. Slimes processing is hindered by the particle size limits of current process equipment and this dissertation explores the potential of coupling selective flocculation with magnetic and gravity separation to improve separation efficiencies. Base case tests (without selective flocculation) were conducted on the SLon-100 (laboratory scale pulsating Wet High Intensity Magnetic Separator) and the laboratory scale Reflux Classifier (RC). The base case tests were conducted to determine the optimal intensity for the SLon-100 and a semi-batch test was done on the RC to determine the effect of increasing water fluidisation rates on the response variables namely Fe concentrate grade, Fe concentrate recovery and separation efficiency. Thereafter selective flocculation conditions were optimised by coupling the process with magnetic separation in order to determine the effect of the operating variables on the response variables mentioned above. A Box-Behnken design was utilised and the ANOVA models developed for the significant response variables were used to optimise the selective flocculation process by simultaneously maximising the response variables whilst minimising the three factors (sodium oleate, paraffin dosage and conditioning time). The optimised selective flocculation conditions were then coupled with the RC in order to compare magnetic and gravity separation with and without selective flocculation. The optimised selective flocculation conditions (1 kg/ton sodium silicate; pH 10; 500g/t sodium oleate; 1431.1g/t paraffin and 4.6 min conditioning time) coupled with magnetic separation showed improved metallurgical performance when compared to the base case test. Selective flocculation coupled with magnetic separation improved the magnetic product Fe grade from 52.28±0.38% to 59.21±0.42% Fe whilst simultaneously improving the separation efficiency from 40±1.46% to 56.8±2.0% and maintaining the Fe concentrate recovery within the 95% confidence limits (69.9% to 72.1%). These results were achieved under laboratory and ideal conditions and may differ from industrial scale results. Inconclusive results were achieved with selective flocculation coupled with the RC and additional testwork is recommendedItem Flotation of non-sulphide PGM ores - Optimization of flotation reagent suite and conditions(2018) Sekgarametso, KatlegoThe aim of this study is to improve the flotation of non-sulphide PGM ores from the Mimosa Mine in the Great Dyke of Zimbabwe by evaluating a variety of collector reagents that have not been tested on such material before and applying a full factorial experimental design to investigate the effects of the main primary collector, co-collector and depressant on PGM recovery and grade. The mineralogical studies by XRD revealed that the non-sulphide PGM ore had substantial amounts of gangue material, comprising of 45% quartz, 21% chabazite and 33% of magnetite. The ICP-OES analysis showed that this particular non-sulphide PGM ore is a low-grade ore with an average 4E head assays of 2.37ppm. In the preliminary flotation stage, three reagent suites made up of (i) a collector, (ii) a co-collector and (iii) a depressant i.e. (SIBX, DTP, M98B); (SIBX, C7133, M98B) and (SIBX, AM810, M98B) respectively were tested. It was observed that (SIBX, AM810, M98B) reagent suite gave the best performance with respect to both recovery and grade of the PGM concentrate from the ore. Attempts were made to optimize the dosage levels of the 3 reagents. The optimization studies revealed that 78.5% Pt and 69.3% Pd can be recovered at grades of 17.90g/t Pt and 9.44g/t Pd respectively. This represents a significant upgrade for the roughing stage from the 1.42g/t Pt and 0.85g/t Pd in the feed. These results were obtained at optimized dosages of 86g/t SIBX and 80g/t AM810, with depressant M98B at 50g/t. The observations from the experiments indicated that recovery of PGEs was on the upward trend as the dosage of hydroxamate was increasing hence the effect of the hydroxamate co-collector was further tested at higher dosages while fixing SIBX at 100g/t. The experiments were carried out using 50g/t, 60g/t, 70g/t and 80g/t hydroxamate (AM810) with the depressant M98B at 50g/t. It was observed that the Pt recovery only increased slightly with increasing hydroxamate (AM810) dosage.Item Measurement and modelling of bubble size in flotation froths(2018) Tshibwabwa, Eric MukendiThe flotation process is widely used for upgrading valuable minerals in the field of mining. Many diverse minerals, including most of the world’s base and precious metals are processed by flotation process. Most valuable products produced by flotation pass through the froth phase of the flotation process. The froth phase has attracted more research in recent times because of its significant role in determining the mineral grade and recovery achieved from a flotation operation. The complex processes that occur in the froth phase – detachment, re-attachment, coalescence of bubbles, and competition for attachment sites, mixing and transport all combine to affect the net transfer of mineral particles into the concentrate. Bubbles are formed at different sizes in the pulp phase and coalesce at different rates and as a result the bubble size distribution varies from point to point in the froth phase. Substantial coalescence gives rise to loss of bubble surface area and hence loss of recovery. Competition for attachment sites gives rise to an increase in grade. No method for measuring the variation of froth bubble size distribution (FBSD) was available until Bhondayi and Moys developed one. The method measures the intrabubble impact distance in the froth using a probe dropped at known height through the froth. The average of these intra bubble impacts was considered to be a proxy for froth bubble size distribution; this was calibrated using FBSD. However the measured in the laboratory using photographs taken through the transparent wall of a laboratory cell. A 31 % of error was found and compared to the photographic method, which indicated that the technique over-estimates the actual froth bubble size distributions. This is due to the use of an average IID (proxy) as an estimate of the bubble. In response to the known of actual froth bubble size distribution FBSD in order to quantify the complex processes in the froth phase, an application of a stereological technique/model was developed and tested to obtain estimates of the actual froth bubble size distribution FBSD in lab flotation and Mintek pilot rougher cells as a function of froth height, frother dosage and superficial gas velocity. The model was first validated for a system of flotation with variable froth height in a transparent Wits lab flotation cell. The two-parameter normal distribution model FBSD was considered to fit the model-predicted intrabubble impact distance distribution IIDDs to measured intrabubble impact distance distribution IIDDs. The model was seen to accurately iii predict the FBSD compared to actual FBSD data obtained from above-mentioned conventional photographic method using a calibration scale attached to the transparent flotation cell wall, wherein the experimental IIDDs were accurately fitted by the model-predicted IIDDs. Similar estimation of froth bubble size distribution were also found with the inversion matrix technique. Secondly, the model was then evaluated for flotation condition with variable frother dosage in the Mintek pilot plant rougher cell. The model was seen to estimate the actual FBSD, wherein the IIDDs were precisely predicted compared to experimental IIDDs. Finally, the model validity was then tested for various systems of flotation conditions with variable superficial gas velocity. The model was seen to estimate the actual FBSD for these cases compared to both model-predicted IIDD and experimental IIDDs. The performance of the present model for these systems of flotation was seen to estimate froth bubble size in froth phase from measured IIDD information. Froth bubble size increases with increasing in froth height, and decreases with increasing in frother dosage and superficial gas velocity. Froth height, frother dosage and superficial gas velocity have a strong effect on froth bubble size distribution.Item Recovery of PGMs from an oxide ore by flotation and leaching(2018) Sefako, Relebohile BasilFroth flotation is the process used in the Platinum Group Metal industry to upgrade the run-of-mine ore for subsequent processes such as smelting and hydrometallurgical PGM refining. The PGM concentrator plants achieve high PGM recoveries (>85%) when treating prestine (unweathered) sulphide ores. However, the depletion of prestine sulphide PGM bearing minerals has triggered interest in exploration of techniques for PGM recovery from near surface oxidised PGM ores. All earlier attempts to process the oxidised PGM ores by conventional flotation methods achieved poor recoveries (typically less than 50 %) hindering the commercial exploitation of these resources. The characterisation of the non-sulphide PGM ore used in this study indicated that the ore is enriched in oxide iron minerals as a result of weathering. In the flotation work, the maximum PGM flotation recoveries achieved using the sulphide co-collector schemes were 55.1% 3E (Pt, Pd and Au). Application of the hydroxamate oxide collector improved the flotation performance to recoveries of 74.7% 3E. The superior PGM recoveries achieved with hydroxamates probably lies in their ability to form complexes with metals such as iron. Hydroxamates co-collectors have been proven to improve recoveries without any adverse effects on performance of primary collectors such as SIBX. In this study the non-sulphide PGM ROM ore was leached directly using different acids. Low PGM extractions were recorded for hydrochloric acid (36.6% Pt and 8.8% Pd) and nitric acid (34.5% Pt and 7.1% Pd). The best leaching results of 48% Pt and 24.5% Pd were obtained using aqua regia solution though it is non-selective. Leaching of ROM ore is generally not preferable as it leads to high reagent consumptions. In this study it was postulated that leaching of low grade flotation concentrate would be preferred. Experiments were conducted to leach the concentrate that had the highest PGM recovery with sulphuric acid in order to target the base metals and further concentrate the PGMs in the residue. The base metal recovery from flotation concentrate using sulphuric acid was only efficient for copper and nickel while poor iron recoveries were achieved.Item Optimisation of reagent addition during flotation of a nickel sulphide ore at the Nkomati Mine concentrator(2017) Kahn, RiyardBatch scale laboratory testwork was conducted to evaluate collector and depressant addition on flotation performance of a nickel sulphide ore. The objectives of the study were to: 1. develop an understanding of the effects of collector and depressant dosage, and its interactive effects, on flotation performance and 2. determine the effect of stage dosing collector and depressant on flotation performance. Testwork was conducted on the Nkomati Main Mineralized zone orebody, a nickel sulphide orebody in the Mpumulanga Province of South Africa consisting of pentlandite, chalcopyrite, pyrrhotite, pyrite and magnesium bearing silicates. Characterisation testwork was conducted, including mineralogy on the major plant streams (by QEMSCAN) and a process survey. The results indicated that there was potential to increase the recovery of coarse pentlandite and that major nickel losses were observed in ultrafine pentlandite. Milling optimisation requires the minimisation of ultrafine generation while ensuring adequate liberation of the course nickel. Stage dosing of collector at nodal points (where more than one stream meets) is currently practiced on the plant, however, its effect had not yet been quantified on the plant or in the laboratory. Stage dosing of depressant is currently practiced on the cleaner flotation stage, however, this too has not been compared to upfront dosage on its own. Significant gangue depression was noted specifically for the cell at which stage dosing was done. The current study would provide an understanding of the current practices with the possibility of offering improvements. The addition of collector progressively improved the hydrophobicity of the sulphide minerals and gangue (with particular emphasis on magnesium bearing gangue), improving recovery significantly. As a result of additional gangue recovery at the higher collector dosages, increased depressant dosages were required to maximise nickel recovery. The collector improved valuable mineral recovery, however, gangue recovery was increased simultaneously, albeit at a reduced rate or in reduced quantities. Furthermore, increased gangue entrainment was evident at higher collector dosages from the increase in water recovery. Excessive depressant addition destabilised the froth phase by the rejection of froth stabilising gangue, which resulted in reduced recovery of the valuable minerals. Therefore, a careful balance must be maintained in order to maximise nickel recovery. Iron recovery was markedly increased at higher reagent dosages, indicative of increased pyrrhotite recovery. Pyrrhotite, although containing nickel, reduces the concentrate grade and may need to be depressed in the latter stages of flotation to ensure the final concentrate specification is achieved. This is an important observation as any improvement in nickel recovery in the roughing stages must be evaluated against the subsequent effect on the cleaning stages. Stage dosing both collector and depressant, individually and collectively, proved to be beneficial by improving the nickel recovery. Stage dosing of both collector and depressant produced higher recoveries than stage dosing of the reagents individually. The time at which the reagent is dosed also proved to have an effect on the performance with an increased dosage in the latter stages providing the highest recovery. The typical recovery by size performance for flotation is characterised by low recovery of fines and coarse with an optimum recovery of an intermediate size fraction. Stage dosing ensures that fine particles are recovered with minimal reagent addition upfront, thereby, coarser particles can be effectively recovered once the high reagent consuming fines are removed. The results have indicated that stage dosing improved the recovery of both coarse and fine particles, whilst reducing the recovery of the intermediate size fraction. Stage dosing can be implemented for two reasons: 1. maximising recovery 2. minimising reagent consumption to achieve the same recovery as upfront dosing A financial evaluation should be conducted to quantify the optimum operating solution. Minimising reagent consumption could be beneficial under conditions of very low commodity prices and excessive reagent costs.Item Process evaluation of column flotation at Ergo(1990) Eves, Jonathan Charles JoshuaThe static Rand gold price has put pressure on the South African gold mines to improve efficiency. Superior metallurgical performance and lower costs attributed to column flotation prompted the construction of a pilot plant (238 millimetres by 10 metres) at ERGO, an Anglo American tailings retreatment plant [Abbreviated Abstract. Open document to view full version]Item Beneficiation of Waterberg Coal(1992) Eroglu, BerrinModern methods of mechanised mining and the necessity for the utilization of total reserves have caused the inclusion of more and more impurities in run of mine coal. This fact, together with the limited supply of naturally clean coal fCI gasification, liquefaction and metallurgical purposes, has made some Iorm l){ beneficiation obligatory at many mines not only in South Africa but also in many other countries. One of the South African Coalfields, Waterherg, contains the continent's largest reserves (approximately 46% of South African known reserves). At the Grootegeluk Coal Mine, approximately 15 m tons of coal per annum are mined by opencast methods. The coal is characterised by containing a high proportion of reactive macerals. The Waterberg Coalfield is currently supplying coal for coke manufacture and middlings for power generation. This coal could also be used for other markets, as Waterberg coal is low in oxygen, contains up to 30% volatile matter. Because it contains 90% vitrinite, it is suitable for direct liquefaction, and possibly coal-water mixtures. However the yield of coal suitable for coking or liquefaction (approx 10% ash) is only 12%, with another 24% of 35% ash coal, currently used for power generation. These yields render mining generally uneconomical if making a simple product. The objective of this project is to ascertain whether the yields of washed coal from the Waterberg Coalfield might be increased by using comminution. Thereafter appropriate beneficiation techniques might be employed on different size fractions. Liberation, float and sink, froth flotation and oil agglomeration processes were examined to identify the best way of treating the coal. Work was carried out on the existing clean coal, middlings and discard fractions. The major objective was to optimise the yield of 10-15% ash coal.The results of the experiment indicate that it is possible to obtain low ash coal from middlings, and middlings from discard for power station. The capital and operating costs for improved new plants are calculated by using available factorised data. The results of experiments on both middlings and discards indicate that yields are significantly higher than those currently obtained, but the cost of obtaining such enhanced yields can be too high for normal commercial application.Item An electrochemical investigation into the floatability of pyrrhotite(1998) Buswell, Andrew MarkImpala's Minerals Processing Plant in the Rustenburg Area, South Africa, uses flotation to beneficiate precious metal bearing ores from the Bushveld Complex. Pyrrhotite is one of the sulphide minerals that is targeted but it is the least amenable to current flotation conditions having the lowest recovery. Electrochemical techniques (mixed potential measurements, cyclic voltammetry and current transient techniques) were used to study the relevant reactions on the surface of pyrrhotite mineral electrodes. Aspects investigated included the oxidation of the mineral in aqueous alkaline solutions, activation by copper sulphate, kinetics of oxygen reduction and the adsorption of isobutyl xanthate. Mixed potential measurements of mineral electrodes were taken in batch flotation test work. In addition a novel qualitative measure of hydrophobicity was investigated. The oxidised surface of pyrrhotite is likely to be covered with iron hydroxides and a sulphur rich sub-lattice. No direct evidence was found for the activation of pyrrhotite by copper sulphate in alkaline solutions. It was shown however that activation could be achieved in mildly acidic media and that the surface remained activated if subsequently exposed to alkaline conditions. When achieved under acidic conditions activation was observed to enhance the degree of interaction between the mineral and the xanthate collector. Also copper sulphate appeared to aid the formation of a more hydrophobic surface (as indicated by the hydrophobicity tests). Copper activation conducted in acidic media did not significantly enhance the kinetics of oxygen reduction, a reaction seen as crucial to the adsorption of xanthate. No evidence was found for the initial chemisorption of xanthate onto the mineral surface. However evidence was found for the oxidation of xanthate to dixanthogen at sufficiently anodic potentials. It Was concluded that the relatively poor flotation performance of pyrrhotite could be combated by minimising the extent of the oxidation, adding reagents as soon as possible before the mineral becomes extensively oxidised and by removing surface hydroxides through lowering the pH during conditioning.Item Reducing the magnesium oxide content in Trojan's nickel final concentrates(2016) Pikinini, SebiaTrojan Nickel Mine in Bindura, Zimbabwe, produces nickel concentrates which, until 2008, were then processed at their smelter operations (Bindura Smelter and Refinery) and the subsequent product sent to the hydrometallurgical plant to produce nickel cathodes. However, due to economic challenges the smelter and hydrometallurgical plant operations were closed down in 2008. Currently, Trojan Mine produces nickel concentrates through flotation which are then sold to Glencore International, in China, for further processing. Since 2002, the MgO (also known as talc) content in the Trojan Nickel Mine final concentrates has increased from around 12% to a peak of 22%. The average MgO content in the concentrates for the year ending in March 2015 was 16.14%. An offtake agreement of sale was made with Glencore International, in China, whereby a penalty is charged for all concentrates with MgO levels greater than 5%. In the year 2015 alone, monthly revenue due to smelter penalties amounted to an estimated total of US$141 000. Higher MgO levels in the concentrates are prevalent when processing low grade ores, with nickel content ranging from 0.65-1.2%. This research focused on reducing the MgO content of the Trojan’s final concentrate to 12%; which was the smelter’s set target while it was still operational. In order to investigate the effect of pH and chemical depressants on the MgO levels in the concentrate, batch flotation tests were carried out at pH 8.95 and 10.2, using several guargum depressants namely: Betamin DZT 245 (standard), Cytec S9349, DLM PDE, DLM RS, and CMC (carboxy methyl cellulose) depressants namely: Depramin 177, 267 and 347, and ND 521, 522 and 523. The concentrates were collected at 1, 5, 15 and 25 minute intervals in order to understand the stage-wise recovery of nickel and MgO minerals. A flotation test, without a depressant, was also carried out in order to understand the kinetics of the gangue minerals. Stage addition of depressants was investigated, by adding another 50g/t dose of the DZT 245 depressant after 1 minute into the flotation test. Collector combination tests using SIPX, SIPX:NC228, SIPX:NC236 and SIPX:PNBX, were also carried out to determine the best reagent suite. To understand the recovery of nickel and MgO in the flotation circuit, a plant survey was carried out, and the particle size distribution (PSD) and assays of collected samples were determined. Flotation tests results indicated that DLM RS and DLM PDE guargum depressants had better selectivity towards MgO and higher nickel recoveries as compared to the Betamin DZT 245 depressant that is currently used in the plant. It is recommended that a plant trial be carried out using the DLM RS depressant, which further reduced the MgO and mass of concentrate recovered by 3.79% and 32% respectively. The stage recovery of MgO for a test carried out without a depressant showed that 57.7% of the MgO was recovered during the first five minutes of the test. Thus, there is need to effectively depress the fast floating MgO during the early stages of the flotation process. Nickel recovery and grade were increased by 2.7% and 2.1% respectively, after adding the second dose of the depressant after 1 minute into the flotation test. The results indicated that the fast floating MgO can depress the valuable mineral if the depressing effect of the depressant is short-lived, which in turn leads to reduced nickel recoveries. Hence, reducing the time between the two stage additions of the depressant in the plant will help further supress the fast floating MgO silicates. It was also noted that at least 60% of the nickel was recovered during the first five minutes of the tests. Hence, reducing the residence time of the rougher flotation bank would reduce MgO recovery into the concentrates without adversely affecting the nickel recoveries. Plant survey results showed that the scavenger bank feed that was deslimed, had less finer MgO particles and MgO content as compared to the rougher bank feed. This indicates that desliming before the coarse flotation process could reduce MgO slimes in the feed, reduce the recovery of MgO due to slime coatings in the final concentrates and the reagent consumption in the bank. Introducing the desliming unit could be beneficial since the desliming cyclones have low installation and operational costs.