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By J. D. Steenkamp, J. Muller

"Excavation of an industrial-scale submerged arc furnace, utilized in the production of silicomanganese, identified two high-wear areas in the refractory lining: the tap-hole, built with cold rammed carbon paste and SiC bricks, and the hearth, built with cold rammed carbon paste. To obtain insight into the potential causes of wear, thermodynamic calculations were conducted on eight sets of data, obtained for seven furnaces on three plants based in South Africa. FACTsage software and associated databases – FACTPS, FToxid, and FSstel – as well as the Mn-Fe-Si-C database of Tang and Olsen (2006), were applied. Theoretical indications are that chemical reaction between carbon refractory and slag, as well as dissolution of carbon and SiC refractory in metal, contributed to tap-hole refractory wear; and dissolution of carbon in metal contributed to hearth refractory wear. IntroductionDuring the excavation of a 48 MVA submerged arc furnace (SAF) used for silicomanganese (SiMn) production, it was found that the taphole and hearth were high refractory wear areas (Steenkamp, 2014). In both these highwear areas, carbon-based cold ramming paste formed the hot face refractory lining (crucible). The wear profile of this SAF is presented in Figure 1.The hot face refractory lining was formed by cold ramming of high-grade carbon ramming material. This material consisted of 50–70% anthracite, 15–25% graphite, 6–12% resin, 2–7% tar, 1–5% clay, and 1–5% alumina. The tap-hole was built using SiC bricks, consisting of 75% SiC, 23.4% Si3N4, 0.3% Fe2O3, 0.3% Al2O3, and 0.2% CaO.In the tap-hole area, the wear of the SiC refractory material was estimated at 0.4 t, and that of the high-grade carbon ramming material, at 1.9 t (Steenkamp, Pistorius, and Tangstad, 2015). The amounts of slag and metal tapped during six years of operation of the tap-hole (September 2007 to April 2013) were estimated at 33 088 t (Steenkamp et al.,2015) and 41 360 t, respectively."

Jan 1, 2016

By M. H. Moys

Accurate models of mill power are becoming more necessary as the design, operational, and control requirements for grinding mills become more stringent. Models of mill power are reviewed briefly, and a new semi-phenomenological model is derived. The conventional torque-arm model for mill power is combined with a model of the tendency of the load in a mill to centrifuge at high speeds, to describe mill power that can be related to several design variables (such as mil geometry, mill speed, and liner design) and control variables (such as load volume, slurry rheology, and mill speed in the case of variable-speed milling). The semiphenomenological nature of the model allows rational conclusions to be reached concerning the effect of various aspects of mill-load behaviour on mill power. The model is shown to be capable of describing reality with much greater accuracy than is achieved by empirical models such as the Bond model.

Jan 1, 1993

By L. Ntlhakana, K. Khoza-Shangase, G. Nelson, I. Maposa

Occupational noise-induced hearing loss (ONIHL) is associated with various risk factors in South African mines. We investigated the association between standard threshold shifts (STSs) and exposure to noise and platinum mine dust (PMD), using demographic data, five years of annual audiometry screening results, and noise and dust exposure data. Miners’ age, sex, percentage hearing loss (PLH), and dust and noise exposure data were gathered, and a linear mixed effects regression model used to predict STS. Average occupational exposure levels to noise and dust were calculated from recorded measurements. A total of 12 692 records were analysed. Most miners were male (89.6%) and more than 50% were younger than 41 years. More than 70% were exposed to > 85 dBA noise and 58% were exposed to 1.5–2.99 mg/m3 PMD. Changes in hearing levels ranged from 8.3 dBHL at baseline (2014/2015) to 10 dBHL in 2016 (STS), with no changes thereafter. The model showed that the estimated effect on STS for males was 27% and 21% higher than for females, for the left and right ear, respectively The estimated effect of age, PLH, noise exposure and years of exposure on STS was < 10% for each variable. There was no statistically significant association between PMD and STS. The combined effects of age, sex, years of exposure to noise, and noise exposure levels, and strength of associations can be used to predict STS for this group of miners. Our findings may be used to measure the efficacy of the mine’s hearing conservation programme.

Aug 1, 2021

By J Taguta, B. McFadzean, P. E. Ngoepe, M. G. Mulaudzi

Computational modelling methods and isothermal titration calorimetry (ITC) were employed to investigate the interaction of a pyrite surface with thiol collectors. Modelling work was conducted in the presence and absence of water. Microcalorimetry was employed to investigate the resultant energy of interaction pyrite as a result of mineral-collector interactions. Sodium ethyl xanthate (SEX), diethyldithiocarbamate (DTC) and diethyl dithiophosphate (DTP) were investigated in this study. The results showed that the calculated surface energies and morphologies for FeS2 showed that (100) and (111) are the most stable surfaces. The morphology also indicated the appearance of the (100) and (111) surface facets and no presence of others. The calculated adsorption energies between thiol collectors and the pyrite surface were always more exothermic than the experimentally determined ones. When comparing different ligand types with the same alkyl chain length, the heat of adsorption of diethyl-DTC was greater than that of ethyl xanthate and diethyl-DTP for pyrite. Experimentally, the diethyl DTP barely adsorbed onto the pyrite surface and this may be due to a different mechanism of action. In computational calculations, water adsorption was found to reduce the reactivity of the Fe atom for the interaction with thiol collectors, and bring the adsorption energies closer to the magnitude of the experimental values. Keywords: Computational modelling, microcalorimetry, experimental and thiol collectors, adsorption energies

Jan 1, 2020

By O. Naiker

Xstrata has modified and improved on the solid state reduction of chromite (SRC) process that was pioneered by Showa Denko in Japan and previously used at Xstrata?s Lydenburg plant. The history of pre-reduction is a long one, with the SRC process being the first and only commercially successful one prior to Premus. Krupp developed the chromite direct reduction and the rotary hearth for the pre-reduction of chromite. However, these processes have not been successfully commercialised. The are a number of significant differences between Xstrata?s patented Premus and the original Consolidated Metallurgical Industries (CMI) process. These include, the use of lower cost anthracite fines as the reductant source together with the use of unactivated bentonite as the binding agent. In addition the use of oxygen as an energy source is used in Premus but not in the original CMI process. There was a fundamental change in the operational philosophy of the process in that while the original CMI process aimed to maximize metallization of the pellets, the Premus process seeks to maximize the energy output from the kiln while still achieving the required efficiencies and hence increasing furnace output. Premus is designed to reduce electrical energy consumption, provide high recoveries of metallic oxides, and utilizes low cost reductant and energy sources such as anthracite and oxygen respectively. Due to the fact that Premus is one of the most technologically advanced and competitive processes used today in the production of ferrochrome it was the process selected for Xstrata?s new ferrochrome facility, Project Lion, that is due to be commissioned in the third quarter of 2006 in the Steelpoort Valley in Mpumulanga, South Africa.

Jan 1, 2007

By E. Kapilashrami, S. Seetharaman

Refractory/metal interactions and associated reactions are significant during casting as they can lead to inclusion formation. High-temperature interactions between aluminosilicate refractories and molten steel were studied in the present investigation using the sessile drop method. The changes in the contact angles between molten iron with well-defined oxygen levels and substrates of silica, mullite (3Al2O3?2SiO2) as well as commercial refractories were measured in the dynamic mode. In all cases, the reaction led to slag formation. The experiments were performed under isothermal conditions at 1823 and 1873 K. A known and constant oxygen partial pressure was imposed above the iron drop, through appropriate CO-CO2-Ar mixtures. For all substrates, the contact angles started decreasing due to the lowering of the surface tension of iron, as oxygen was imposed on to the system. When a critical level of oxygen was reached, slag started forming at the drop/substrate interface and at this stage, the contact angle dropped suddenly. Subsequent behaviour of silica and mullite was observed to be different, with SiO2showing a non-wetting behaviour. In the case of SiO2substrate, there were also deep erosions along the periphery of the drops and there was slag accumulation at the sides of the metal drop, probably due to Marangoni flow and/or SiO(g) formation. SEM and EDS analysis of the substrates revealed that the formed slag was fayalite. The thermodynamic criteria for slag formation and possible mechanisms of the reaction are discussed in the light of the experiments. Key words: silica, mullite, fayalite, corundum, slag formation, sessile drop.

Jan 1, 2004

By A. Singh

The control of flotation plants is by no means trivial. Flotation is one of the most interactive minerals? processing operations. The interactive nature of flotation circuits results in oscillations of levels and hence of grades and recoveries. Also, poor level setpoint tracking and disturbances have negative influences on grades and recoveries. The result is an unstable plant with limited scope for optimization. An advanced controller, FloatStarTM, has been developed to provide tight control and to reduce the amount of disturbances in flotation levels. It has been shown on a number of plants that the system results in faster and better setpoint tracking, better regulatory control and faster settling times after start-ups and disturbances. The successful stabilization provided by FloatStar has made it easier to investigate the area of optimization on flotation circuits. Mintek has identified the need for a more meaningful set of setpoints or aeration rates to use as a basis for control than levels and have advocated flow control by looking after residence times. When the feed conditions change, the ideal solution would be an algorithm or strategy that finds the new optimum conditions rather than just having an unguided hunt from the previous point. A multivariable flow-optimizing algorithm, a new module in FloatStar, has been developed to calculate optimum level setpoints and/or aeration rates that will aim to optimize the residence times, mass pulls and circulating loads within a flotation circuit. Keywords: flotation, regulatory, FloatStar, start-up, disturbance, residence times, circulating loads

Jan 1, 2003

By A. Amiel

Tarkwa Gold Mine is owned by Gold Fields Ghana, IAMGOLD and the Government of the Republic of Ghana in a 71%, 19% and 10% ratio. The mine is operated by Gold Fields Ghana. This large, low-grade, open pit, gold mining operation is located in Tarkwa, in the Western Region of Ghana, approximately 320 kilometres by road, west of Accra, the capital of Ghana, West Africa. In the Financial year ended June 2007, the operation mined a total of 108 million tons of which 22 million tons was ore. The mined ore is processed through three gyratory crushers supplying two heap leach pads (north and south) and a carbon-in-leach (CIL) plant. Before the 6 Sigma project commenced, the current fragmentation was resulting in sub-optimal feed size to these crushers. It was decided to use 6 Sigma methodology to improve the fragmentation. A defect was defined as a fragmentation analysis indicating less than 95% passing 750 mm. The objective was to improve fragmentation in order to achieve a constant throughput to the north heap leach crusher of 1500 t/h, a 5% increase in pit loading efficiency and a 50% decrease in rock breaker hours, equipment damage costs and in-pit sheeting costs respectively. However, this was not to be achieved at the expense of grade control as Tarkwa is exploiting a low grade, high volume deposit and therefore ore dilution must be kept to a minimum. This paper is intended to take the reader through the various 6 Sigma phases of the project (define?measure?analyse ?improve?control), which have resulted in a 10% reduction in loading times, of which half has been attributed to improved fragmentation, as well as a substantial increase in blasted stocks. The focus has been on both drilling and blasting parameters as both processes had to be taken into account. These parameters are technical but personnel factors were also identified and dealt with as they arose. In order to measure the impact of any change in the process parameters on the fragmentation size distribution, cameras have been acquired. One has been installed looking into the bowl of the north crusher and the other one was dedicated to field observation, of individual blasts. Both are used for ongoing fragmentation analysis. At Tarkwa, mining efficiency is strongly dependent on good fragmentation as well as the presence of unblasted ?toes? on the pit floors. Crusher throughput depends on an acceptable ore fragment size. Compliance with plan for actual hole depths was temporarily monitored by manual hole dipping, but this is now being replaced with a fully automated drill data management system developed by Sandvik and Modular Mining. The Six Sigma methodologies, using control charts based on data collected in the field, allowed the mine to improve methods for the control of blasted material. It also enforced the implementation of a quality control programme encompassing the entire drill and blast process.

Jan 1, 2008

By Ike Madingoana, Stephan van Zyl, Antoine F. Mulaba-Bafubiandi

"The study presents a control strategy implemented at the Nkomati nickel mine flotation plant for process stabilization and improvement. The control strategy involves flotation concentrate volumetric flow rate stabilization and the flotation cell mass pull control. The flotation concentrate volumetric flow rate is operated at a fixed set point. The initial control strategy involved varying the flow rate to maintain the concentrate tank level thereby reducing pump cavitation. The flotation cells mass pull control comprises air rate manipulation into the cells with the exception of the recleaners. The air rate into cells is automated to adjust based on the concentrate tank level variation. The control strategy was implemented on the roughers, cleaners and the recleaner circuit. The recleaner cells air flow rate is adjusted based on the MgO content in the concentrate. The stabilization of the flotation concentrate volumetric flow rate and the mass pull by aerate automation will increase the scope of optimization for further process improvement. The immediate benefit of the control strategy implementation was plant stability. Disturbance to the flotation cells downstream was minimized due to stable flow rate, and the air addition manipulation ensured that the concentrate tank levels are maintained, minimizing pump cavitation. INTRODUCTION Froth flotation is a challenging process because it is influenced by a large number of factors, including the flotation feed characteristics. One of the ore types mined at Nkomati nickel mine is the main mineralized zone (MMZ). The MMZ ore is hosted in the Uikomst complex, a Ni-Cu-PGE-Cr- mineralized, layered basic to ultra-basic intrusion which hosts commercially viable magmatic base-metal sulphides (De Waal et al., 2001). The MMZ ore is a heterogeneous sulphide mineral resource with a varying sulphide mode of occurrence and texture. The beneficiation of the complex sulphide ore is challenging due to the presence of impurities in the ore. The impurities report to the flotation final concentrate and penalties are levied if the impurities are beyond the contractual limits."

Jan 1, 2015

By B. P. Watson

The discussion written by Dr de Frey suggests that mine design should be based on closely spaced stability pillars. However, this concept leads to low extraction ratios at depth. The concept of `crush&apos; pillars in conjunction with more widely, appropriately spaced stability pillars allows for significantly larger extraction ratios, particularly at depth. It should be noted that the rock conditions in many areas of South African platinum and gold mines are good, allowing significant spans to be mined safely between stability pillars. `Crush&apos; pillars are cut small enough to fail stably near the face under stiff loading conditions, and their residual strength is intended to support about 10 m rock height between stability pillars. While this support resistance cannot be achieved by conventional support, crush pillars have been proven to have the required capacity for this purpose. The concept of leaving smaller crush pillars with appropriately positioned stability pillars has been used successfully for more than 30 years in areas of South African platinum and gold mines.

Jan 1, 2011

By N. D. L. Burger

An effective prototype face support system to minimize rockfalls in rockfall-prone mines was developed and patented during a project sponsored by the Mine Health and Safety Council (MHSC). An experimental development model (XDM) roof support unit was developed and evaluated. Mines with rockfall problems in the face areas can use the roof support system to reduce fatalities and injuries in the face area of stopes. The roof support system consists of two similar support units connected to each other via two crank mechanisms. Each unit consists of a headboard supported by a ?wishbone? structure (top and bottom leg). In the first design a threaded bar with struts similar to a scissors jack keeps the legs apart. (In the final version that was field trialed, the threaded bar mechanism was replaced with a hydraulic cylinder). All the components are manufactured from steel. To move the system forward the first support unit is collapsed, the second remaining in the loaded position. The first unit then hangs from the second unit to which it is connected. The first collapsed unit is then manually cranked forward and prestressed. If the required position is still not achieved, the other unit is released and moved forward. The specification for the system was determined and presented at a workshop with industry participants. Different concepts were developed and evaluated against the system specifications. A technology demonstrator was then developed and tested on surface. The technology demonstrator development process included detail design, building and testing of components and subsystems, design reviews and the building and commissioning of the technology demonstrator. The testing of the technology demonstrator was done in a 500-ton hydraulic press, in a mock-up stope and underground. A risk analysis, in which technical, logistical and economic aspects were assessed, was done to determine the critical areas of the system. During the next phase of the project working prototypes were developed for underground tests and evaluation. This process indicated that it was necessary to adjust the system specifications and a redesign was called for. In an iterative process test units were built for purposes of evaluation, specification verification and field trialing. After the successful conclusion of the trials the equipping of a complete production stope started for purposes of integrating the system into the current mining process.

Jan 1, 2008

By M. D. Mwema

Copper oxide ores deposits in the Katanga province of the Democratic Republic of Congo (DRC) contain at variable level cobalt oxide minerals. In contrast to copper oxide minerals which readily dissolve in the sulphuric acid solution, the cobalt oxide bearings are difficult to dissolve, especially if cobalt is present in the third oxidation state (Co2O3). This species become soluble after reduction of Co+3in Co+2. In practice, the use of reducing agent is therefore necessary in order to transform the insoluble minerals into soluble cobalt sulphate species. The following reducing agents are currently used at the Shituru plant: ferrous ions, which are present in leach solutions, pulverised copper or sodium meta bisulphite (MBS). Due to the high consumption of reducing agents (0,8 t MBS/t of solubilized Co; 1,2 t Cu/t-Co), cobalt production is expensive. The use of sulphur dioxide as reducing agent was then investigated in order to reduce operating cost. This reagent also presents the advantage of sulphuric acid production during the leaching process. It has been added in the leaching vessel (tank). From the laboratory and industrial tests, selective leaching of cobalt (78%) has been performed. The cobalt leaching kinetic depend on sulphur dioxide dosage. Dissolution of copper, iron and manganese remained very low. No emanation of SO2 gas was observed during the process.

Jan 1, 2002

By A E. Paterson

General aluminium corrosion charts based on 1100 alloy listing product and temperature have been developed by AFSA, mainly from available sources. The categories of corrosion expectation are fairly wide. As a result the charts are more useful in indicating what is not suited as distinct from what is. The question of how to develop useful information is assessed. Having discarded improbable options one is left with the need to refine the possible. Within each alloy group corrosion potential differs. From an engineering application point of view, the corrosion resisting marine alloys, the 5xxx series show increasing corrosion as the percentage of Magnesium increases. An important issue is that as the welding filler used for 5xxx series is Mg rich (to avoid hot cracking), the filler is likely to show the weakest resistance. In addition, when subjected to fatigue loading, the preference for critical joints is to dress the joint to the level of the base material. Overheating of the weld extends the HAZ zone and weakens the material as Mg moves to the grain boundaries. Indicator chart refinement approaches suited to aluminium use in commercial transport are discussed.

Jan 1, 2006

By D. Tesh

The mining industry worldwide is facing a tremendous shortage of minerals engineers in all fields of specialization. For instance, in Australia the skills shortage in the mining industry has been identified as one of the top risks facing the mining industry. In Namibia most minerals engineers employed are expatriates, with some being Namibians who studied abroad. The minerals engineering programmes at the Polytechnic of Namibia are still in their infancy. These programmes were designed to meet the mining industry skills needs. Being young has its advantages in that lessons can be learnt from older minerals education institutions that went through similar challenges. However, this does not imply just copying and implementing their approaches, since the context differs and, to ensure sustainability of the minerals education programmes, curricula have to be customized to the local context. This paper reviews the interventions made by the Polytechnic of Namibia in order to ensure the sustainability of its minerals education programmes. The methodology consisted of an extensive literature review, a status quo analysis of the mining and process engineering department, identifying the gaps between the current state and desired state, and mapping the goals and strategic actions required to progress to the desired state. The following factors were identified as major threats to the sustainability of minerals education at the Polytechnic of Namibia: the quality of students, low student enrolment rates, low pass rates in science and mathematics, shortage of academic staff, dwindling government funding, and limited involvement of the mining industry. The major outcomes of this research comprise the detailed strategic actions employed by the Polytechnic of Namibia to address the threats to sustainability of its minerals education programmes.

Jan 1, 2014

By D. Yu, Y. Thibault, D. Paktunc

"Prereduction and flux-aided direct reduction of chromite provide significant advantages in reducing energy consumption and greenhouse gas emissions during ferrochrome production. In this investigation, a comparative evaluation of the influences of several alkali fluxes was carried out based on experimental observations supplemented by advanced material characterization and thermodynamic predictions. Direct reduction of a chromite ore with alkali fluxes at 1300°C for 1 hour produced (Cr,Fe)7C3 type alloys with Cr/Fe mass ratios from 0.7 to 2.3. Among the alkali fluxes, reduction aided by NaOH resulted in a high degree (85%) of Cr metallization with the ferrochrome alloy being Cr4.2–4.6Fe2.4–2.8C3. The formation of liquid slag, which facilitated Cr metallization, was limited by the formation of NaAlO2 between 800 and 1300°C. This, in turn, restricted the collection and transport of the charged ionic Cr species (i.e. O2–) to graphite particles. Under the conditions studied, ferrochrome particles were often small and largely unliberated, which would make the physical recovery of ferrochrome challenging. At 1400°C, the amount of liquid slag increased, enabling the growth of alloy particles. Direct reduction of chromite aided by NaOH is promising as an alternative technology to conventional flux-based smelting in electric arc furnaces. IntroductionConventional smelting processes are energyintensive (Riekkola-Vanhanen, 1999) with the energy requirements greater than 4 MWh/t (Naiker and Riley, 2006; Beukes, van Zyl, and Neizel, 2015), and the greenhouse gas emissions can be significant, exceeding 10.5 t CO2 per ton Cr in ferrochrome produced (International Chromium Development Association, 2016). These include emissions occurring on-site (smelter), emissions due to electricity production, and emissions due to upstream processes. Overall energy requirements are influenced by the degree of prereduction and smelter charge temperature. Significant reduction in electricity consumption can be realized if the charge is prereduced before feeding to the submerged arc furnace (Niayesh and Fletcher, 1986). This consideration has led to the development of several prereduction technologies, with the most important example being the Premus process. In comparison to the conventional smelting processes, this prereduction process lowers the overall energy consumption and greenhouse gas emissions by about one-third (Naiker, 2007)."

Dec 1, 2018

By M. Khodadazadeh, K. G. van den Boogaart, T. Leißner, L. Tusa, J. Gutzmer, R. Gloaguen, M. Kern

Ore sorting is a technology that is increasingly used to process primary raw materials. Timeconsuming and expensive empirical state-of-the-art test work is carried out to assess whether the use of ore sorting for the enrichment of a particular ore makes technical and economic sense. With the innovative simulation-based approach presented here, it is possible to direct the selection of a suitable sensor based on quantitative mineralogical and textural data, thus avoiding much of the empirical studies. Required data can be collected quickly and cost-effectively using available methods of automated mineralogy. The obtained parameters such as mineral grain size distribution, modal mineralogy, mineral area and mineral density distribution have been utilised in this study to simulate the prospects of success of ore sorting, applying different types of sensors. Empirical tests with commercially available sensor systems have been conducted to experimentally validate the predictions of the simulations. The estimation of the target mineral grade can be further optimised by the use of machine learning algorithms for the integration of automated mineralogy data and sensor data. The approach can easily be adapted to other types of raw materials and thus has great potential to become a key technology for the optimisation of processing experiments. Keywords: Automated mineralogy, hyperspectral imaging, machine learning, ore sorting

Jan 1, 2020

By Karl-H. Daum

Over many decades, numerous feasibility studies have demonstrated, that the production of sulphuric acid remains the most viable option of sulphur recovery from smelter off gas and abatement of SO2 emissions to the atmosphere. This is particularly more pronounced as smaller but more concentrated off-gas flows are to be treated from smelters, and enhanced sulphuric acid processes be-come available. Off-gas handling systems represent a significant capital and operating cost burden to the metallurgical operation. Modern pyrometallurgical smelter processes for sulphide ores based on the use of oxygen-enriched air, produce relatively small off-gas flows with high SO2 concentrations in the smelter gas of 30?60 %-vol. of SO2. This is a prerequisite for substantial cost reductions in the smelter off-gas handling and treatment system. As an alternative to sulphuric acid production, numerous scrubbing concepts with alkali or dual-alkali combinations as well as organic absorbents have been proposed. Also the reduction of the SO2 to elemental sulphur has frequently been studied. Very few of those alternative processes have been built in industrial scale, but all were generally characterized by none sustainable operation due to cost reasons, problems with issues related to chemicals used and by-products or poor availability. Thus the traditional concept of converting the SO2 to sulphuric acid is most common, although regarded as un-economic, but is at least a proven, environmentally sustainable and reliable way of sulphur gas processing. With the high acid price levels these days, operating companies are even able to generate significant revenue with their otherwise ?fatal? acid.

Jan 1, 2009

By D. Skelton

The Renard kimberlite cluster is located in the Monts Otish region of Quebec, Canada. A Canadian National Instrument (NI) 43-101 compliant resource statement for the Renard kimberlites 2, 3, 4 and 9 and the Lynx kimberlite dyke was first issued in 2008 followed by a Preliminary Economic Assessment and the development of a conceptual mine plan. Following a successful drill program in 2009 that greatly expanded the amount of kimberlite in Renard 2, a revised resource statement was issued in December 2009 comprising 23 million carats of Indicated Mineral Resources and 13 million carats of Inferred Mineral Resources, a threefold increase on the previous estimate. The project is currently the focus of a second Preliminary Economic Assessment and it is expected that a full feasibility and permitting program will commence in 2010 leading to the creation of Quebec&apos;s first diamond mine by 2013.

Jan 1, 2010

By C. Wade, A. Swart, M. Dworzanowski

AngloGold Ashanti operate a pyrite flotation plant at their East Gold Plant, Acid Plant and Flotation Plant (EGAF) metallurgical complex within their Vaal River operations in the North West province in South Africa. The flotation plant feed is reclaimed pyrite rich tailings and the concentrate produced is used as the source of sulphur for the EGAF acid plant. The sulphuric acid produced is in turn used for the recovery of uranium at the Vaal River South uranium plant. The original flotation plant built in 1960 was based on the use of many small volume flotation cells and was becoming very expensive to operate and its age severely limited further operation. A decision was taken in 2005 to replace this plant with a new one based on fewer, larger volume flotation cells. This paper describes the benefits of using larger volume flotation cells in this new flotation plant when compared to the smaller volume flotation cells in the old plant. The benefits of the new plant include a smaller plant footprint, fewer flotation cells, fewer pumps, no distribution boxes, simpler piping, a greater flotation volume and a reduction in installed power. The saving in capital cost could not be quantified because of the age of the old plant. The savings in operating cost will arise from fewer operators, less maintenance and reduced power consumption. The new plant has not been in operation long enough to quantify the benefits in operating cost and metallurgical performance. However, recovery improvements can be expected given the greater flotation volume (increased residence time) and the installation of larger, more efficient flotation cells.

Jan 1, 2007

By R. A. Wildenboer, R. F. Sandenbergh

Rare earth elements (REE) are present at concentrations of approximately 0.36 % in phosphogypsum stacks located in Phalaborwa, South Africa. The REE are present in the phosphogypsum in solid solution in the gypsum and as fluorine and aluminium rich precipitates. Recovery of the REE offers opportunities for valorization and reducing the environmental impact of the phosphogypsum waste. This paper presents the results of investigations into the leaching behaviour of rare earth elements and other impurities from phosphogypsum using aqueous sulfuric acid. The effects of acid concentration, temperature, residence time, and the influence of impurities were evaluated for their impact on the solubility of rare earth elements in a sulfate medium. It was found that a significant fraction of the REE can be leached from the phosphogypsum in a range of conditions much less severe than traditional hard rock rare earth processing conditions. The REE were leached at a sulfuric acid concentration of 110 g/l, temperature of 30°C – 40°C and a residence time of approximately 8 hours.

Nov 6, 2024