3. Electronic Theses and Dissertations (ETDs) - All submissions
Permanent URI for this communityhttps://wiredspace.wits.ac.za/handle/10539/45
Browse
9 results
Search Results
Item Techno-economic optimisation of mining cut for Zimbabwe Platinum Mines' steeply dipping tabular orebody(2019) Wailesi, MaidaZimbabwe Platinum Mines (Zimplats) extracts Platinum Group Metals (PGMs) in Zimbabwe along the Great Dyke. The orebody being mined at Zimplats is tabular, synclinal in nature and plunges gently at approximately 1.5⁰ towards the north. Layers of igneous rocks within the deposit dip at 9° to 20° near the limbs and flatten near the centre to form a flat-lying floor. Mineral Resources at the mine are classified into ‘flats’ (Mineral Resources dipping shallower than 9⁰) and ‘steeps’ (Mineral Resources dipping at an angle > 9⁰). Current mining operations are limited to the ‘flats’ portion of the deposit but as mining progresses towards the north, the ratio of ‘steeps’ to ‘flats’ increases. The current mining layout design was based on the ‘flats’ with a mining cut of 2.5m. In order to exploit the steep dipping orebody limbs, modification of the current mining layout was done in 2016 and trials were conducted on the new design. Results from the trials showed that the design resulted in a lower actual head grade than planned which was attributed to a sub-optimal mining cut. This research study was done to determine an optimum mining cut at Zimplats mine for the ‘steeps’ portion of the deposit, focussing on the Mineral Resources dipping between 9° to 14° to create optimum value from extracting the steep Mineral Resources. The study was carried out at P4 mine and two blocks out of the available six blocks were randomly selected for the purpose of this study. In the optimisation process, six mining cuts were selected, 2.00m, 2.10m, 2.20m, 2.30m, 2.40m and 2.50m. Mine designs were generated for each mining cut in Vulcan 3D software and Mineral Reserve calculations were done for each option. Mining schedules were generated through the application of the Vulcan Gantt Scheduler. A techno-economic evaluation was done on the six mining cuts focusing on, Mineral Reserve tonnage, grade, extraction ratio, cost and revenue generated. A discounted cash flow (DCF) analysis was done which incorporated all the evaluation parameters to determine the optimum mining cut which yields the highest Net Present Value (NPV). A sensitivity analysis was done to determine the reliability of the chosen optimum mining cut. It was found that the optimum mining cut varied depending with the optimisation criteria. However, from the DCF analysis, the 2.10m mining cut option produced the highest NPV of US$95 million among the other possible mining cuts. Therefore, the research study concludes that the optimum mining cut for the ‘steeps’ Mineral Resources at Zimplats mines is 2.10m. It is being recommended that the company adopts 2.10m for the exploitation of its ‘steeps’ resources. It is also recommended that the company should design a mining method for the ‘steeps’ resources with dip angles greater than 14° for maximum Mineral Resource utilisation.Item 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 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 Investigation of the joint comminution and leaching process for a gold ore: an attainable region approach(2016) Hlabangana, NkosikhonaComminution and leaching unit processes play a major role in extracting valuable minerals from ore. Most of the research reported in the literature has focused on optimising individual unit operations rather than on integrating the whole process. This thesis develops an integrated approach to mineral processing systems and flow sheets and is intended to create a methodology for process synthesis that can be applied throughout the extractive metallurgical industry. This could lead to improved efficiency in the overall process by obtaining optimum recovery and, most important, a reduction in energy and material costs. In order to illustrate the methodology a particular example was chosen, namely optimizing the joint comminution and leaching of a particular gold ore. In this investigation laboratory scale grinding and leaching profiles for a gold feed sample (1700–850 μm) were measured. In a laboratory mill various combinations of grinding media, filling level and ball size were investigated, and of the three ball sizes used (10, 20 and 30mm) breakage was most pronounced for the 20 mm. Thus for instance it was also established that when using a higher filling ( =30%) and a ball size of 30 mm, more energy was consumed but less liberation occurred, thus a lower amount of gold was extracted during a 24-hour leaching period. Finally, the breakage kinetics of the gold ore was looked at. Using a standard population model the breakage and selection function parameters were successfully calculated. An investigation into the dissolution kinetics of gold ore in a solution of NaCN was also done. These were found to depend on the stirring rate, reaction temperature, particle diameter and the concentration of the leachant. The rate increased with the stirring speed, reaction temperature and leachant concentration, but decreased when the particle size was greater. The activation energy for the dissolution was estimated at about 3 kcal/mol. Furthermore, the linear relationship between the rate constant and the reciprocal of the square of the particle size is a strong indication that the gold dissolution process is diffusion-controlled. The experimental results were well-fitted to a shrinking core model. In attempting to understand the results, the researcher carried out a number of experiments that involved an investigation into the relationship between comminution and leaching in terms of energy usage and particle size, the former to establish the most efficient application of energy, and the latter to identify the degree of fineness that would ensure optimal recovery. The Attainable Region (AR) method was then used to establish ways of finding the leaching and milling times required to achieve minimum cost (maximise profit). No work on utilizing the AR technique to minimise the cost of milling and leaching on a real industrial ore has previously been published. The investigation aims to show how the AR technique can be used to develop ways of optimising an industrial process that includes milling and leaching. The experimental results were used to show how this method could be successfully applied to identifying opportunities for higher efficiency when performing these operations. The approach however is general and could in principle be used for any two or more unit operations in determining how the product from one unit should be prepared to feed to the next unit so as to optimize the overall process.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 Investigating particle size segregation in a batch jig(2016) Silwamba, MarthiasParticle size and size range are among the characteristics that affect the segregation of particles in a jig hence they affect the separation efficiency. The effects of these variables on segregation of particles are not fully understood. This work aimed at contributing to knowledge in this area. To better understand how particle size and size range influence segregation, tests were conducted in which the effects of the density and shape of the particles on segregation were minimized by using as the feed material spherical glass beads of uniform shape and density. Batch experiments of two components systems of various particle sizes were conducted under the same set of jigging conditions: the jigging frequency and jigging time were respectively maintained at 60 cycles per minute and 999 seconds (16.65 minutes). The effect of these operating conditions on segregation was not investigated. At the end of each test run, the jig bed was split into horizontal slices and the composition of each slice was determined. The experimental results showed that below a particle size ratio of 1.50:1, the driving force for the segregation of particles, i.e. the particle size difference, was small hence a low degree of segregation was obtained. The degree of segregation increased above this ratio. However, above the size ratio of 2.00:1, interstitial trickling occurred. With the smaller particles tested (8, 6 and 4mm) poor segregation was observed when the size ratios were of 1.50:1 or less along with what is believed to have been remixing due to convective currents within the jig chamber. It was found that the particle size range had a more pronounced effect on size segregation than the particle size. From the results, it can be said that above a size ratio of about 1.50:1, size segregation is very pronounced. This suggests that density separations of real ores, where both the density and size of particles vary, would be impaired if the particle size range of the material fed to the jig exceeds this ratio. However, this needs further confirmation by testing multiple component systems.Item Optimization of dense medium cyclone plant for the beneficiation of low grade iron ore with associated high proportion of near-density material at Sishen Iron Ore Mine(2016) Tom, PhakamileThe research report is premised on three aspects which are critical in the heavy mineral beneficiation. These aspects are classified as (i) understanding the densimetric profile of the available ore body, (ii) understanding the properties of the heavy medium utilised at the plant to beneficiate the ore, and (iii) the automation and modelling of the processing plant in order to maximise plant efficiency. Ore characterisation is mainly focused on understanding the densimetric profile of the ore body, in order to determine the probability of producing a saleable product as well as predicting the expected yields and quality. This is done to utilise the endowment entrusted upon the operating entity by the government and shareholders to treat the mineral resource to its full potential. Understanding of the beneficiation potential of the ore body will assist the mine planning and processing plant to optimise the product tons and quality. This will ensure the marketing plans are in accordance with the expected product as beneficiation will vary depending on the mining block reserves. The mining blocks have potential to produce varying product grades with different recoveries. Ore characterisation was conducted on the GR80 mining block, low-grade stockpiles (i.e. C-grade ore reserves & Jig discard and dense medium separation (DMS) run-of-mine (ROM) material. The GR80 material was characterised as having low proportion of near-density material and would be easy to beneficiate as well as produce high volumes of high grade product. Furthermore, it was revealed that the 2014 DMS ROM had an increased proportion of low-density material; however this material was also had low proportion of near-density material. The low-grade stockpiles was characterised by high proportion of near density material, which necessitate the beneficiation process to operate at high density in excess of 3.8 t/m3. Maintaining a higher operating density requires more dense medium which leads to viscosity problems and impact performance. The characterisation of the FeSi medium was imperative to understand its behaviour and potential influence on beneficiation of low-grade stockpiles and mining blocks with elevated proportion of near-density material. As the proportion of near-density waste material increases in the run-of-mine (ROM), it is necessary to beneficiate the material at elevated operating medium densities. However, when cyclones are operated at high densities, the negative influence of the medium viscosity becomes more apparent and thus influences the separation efficiency. Heavy medium, ferrosilicon (FeSi) characterisation looked at identifying the effects of viscosity on the FeSi stability and whether there would be a need for a viscosity modifier. Thus, the importance of controlling the stability, viscosity, and density of the medium cannot be under-estimated and can very often override the improvements attainable through better designs of cyclones. Furthermore, the slurry mixture of the heavy medium utilised for the purpose of dense medium separation should be non-detrimental to the effectiveness of separation in the DMS Fine cyclone plant. Medium characterisation showed that removal of ultra-fines leads to unstable media as indicated by faster settling rates. This would result in medium segregation in the beneficiation cyclone thereby leading to unacceptable high density differential which will negatively impact the cut-point shift and cause high yield losses to waste. The overall control of the metallurgical processes at Sishen’s Cyclone Plant is still done on manually and thus operation still varies from person-to-person and/or from shift-to-shift. This result in some of the process data and trends not being available online as well as being captured inaccurately. Furthermore, this negatively affects the traceability and reproducibility of the production metallurgical key performance indicators (KPI’s) as well as process stability and efficiency. It has been demonstrated that real-time online measurements are crucial to maintaining processing plant stability and efficiency thereby ensuring that the final product grade and its value is not eroded. Modelling and automation of the key metallurgical parameters for the cyclone plant circuit was achieved by installation of appropriate instrumentation and interlocking to the programmable logic control (PLC). This allowed for the control of the correct medium sump level, cyclone inlet pressure, medium-to-ore ratio as well as online monitoring of density differential as “proxy” for medium rheological characteristics. The benefit of modelling and simulation allows the virtual investigation and optimisation of the processing plant efficiency as well as analysis of the impact of varying ore characteristics, throughput variations and changing operating parameters. Therefore it is imperative that all cyclone operating modules are operated at the same efficiency which can be achieved by optimized process through proper automation and monitoring, thereby improving the total plant profitability. Keywords: dense medium separation; densimetric profile; dynamic modelling; FeSi rheology; iron-ore beneficiation; process automation; process control.Item Beneficiation of fine ores using the Longi wet high magnetic separator(2014) Makhula, Mpho JohannaMagnetic separation has been used since 1955 for processing a variety of minerals from iron ore in steel production to the desulphurisation of coal. The accumulation of such fines and slimes during mining operations and the increasing global demand for quality products motivated the use of the semi-continuous pilot wet high intensity magnetic separator (WHIMS) introduced in 2008. Its unique features were considered to be favourable for beneficiating a Sishen low grade hematite-rich ore and an ash-rich Witbank coal. An automated Mineral Liberation Analyser (MLA) was used to characterise the size distribution of the hematite-rich material thereby providing an estimated grade at the same time. This type of analysis characterises mineral types in terms of particle size and elemental composition, specific density, weight percent, area of particle, particle shape, circularity and equivalent circle diameter. However, for this research study only size and elemental composition were considered. The application of the Longi LGS 500 WHIMS for beneficiating a low grade South African iron ore material was investigated by determining the effects of changing the operating parameters of pulp solids, magnetic field intensity and the pulsation frequency. This was followed by a 33 full factorial design which consisted of twenty seven (27) test matrix, with mass yield of concentrate and Fe grade selected as the main responses to the changing of the parameters. The results obtained were validated using the analysis of variance (ANOY A) and the mathematical model, which showed the variables as being significant to the investigation process, thus rejecting the null hypothesis. The significance of the variables was in the order of magnetic field intensity followed by pulsation frequency and lastly the percentage pulp solids. The model predictions and actual data were in good agreement, reporting regression coefficients ranging between 0.83 and 0.94. It was shown that a single stage magnetic separation has the potential to produce a 55% Fe product.