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By T. De Lange

Extensions to the beneficiation plant at the Grootegeluk Coal Mine are currently being planned in order to provide power station coal for Escom's Matimba Power Station, presently under construction. A simulation model was developed on the Olivetti M-24 microcomputer to facilitate the development of the envisaged flowsheet. The computer program simulates the heavy-medium separation and screening unit operations. The model is based on the construction of the Tromp partition curve described by an arctangent function, allowing ideal and non-ideal separations. The paper discusses the determination of the parameters required by the arctangent curve from standard process parameters such as Epm, Wolf cut-point, Tromp cutpoint and screening efficiency. The accuracy and shortcomings of the model are discussed, while an overview of the application of the model to evaluate borehole and bulk sample analyses is also given. It is concluded that the model is an invaluable aid to general flowsheet development.

Jan 1, 1987

By C. J. Fauconnier

The main purpose of the paper is to determine the contribution of fluctuations in barometric pressure to flammable gas explosions in South African mines. For this purpose, the barometric-pressure patterns preceding each of the 59 major explosions in South African gold and coal mines over the past 20 years were analysed. It is concluded that diurnal fluctuations in pressure do not have a major influence on the accumulation of gas in the underground workings, but that pressure drops associated with cyclonic weather systems moving across the country over a number of days are the major factor contributing to gas explosions in mines. To illustrate some of the points made in the paper, the Hlobane Colliery explosion of September 1983 is used as a case study. In the final section, several general conclusions are drawn and recommendations put forward as to how mines should recognize and react to meteorological phenomena that may lead to an increased probability of gas explosions in underground workings.

Jan 1, 1992

By M. J. Deats

INTRODUCTION To prepare the reader adequately, reference should be made to the paper 'The Pioneering of Fully Mechanized Longwall Coal Mining in South Africa' by R. T. Naude and M. J. Deats, published in the February, 1967, edition of the Journal. At the time mechanized longwalling had been proved operationally feasible with encouraging results since its inception in South Africa in mid 1965. Subsequently, further top seam panels were successfully extracted with improved productivity and with some modification to equipment. In 1968, after an overseas visit by mine officials, equipment specific to suit local conditions was acquired and lower seam trials commenced. Despite initial obstacles and difficulties the first panel was successfully mined. Modification and improvements to equipment and techniques resulted in the second lower seam panel being extracted economically during 1969-1970 and with productivity exceeding even that of top seam faces. During extraction of the second lower seam panel, gradual and complete surface subsidence occurred over the panel without affecting face operations. Rock mechanics investigations had originally indicated some uncertainty regarding adverse dolerite sill behaviour which could have caused longwall mining to be hazardous. All such doubts have now been removed. RESUMÉ AS AT FEBRUARY, 1967 Mechanized longwalling was introduced in 1965 on an experimental basis in an effort to improve productivity in the newer area of the mine and to increase the quality of the comparatively dirty run of mine feed to the washing plant. In addition advantages could be predicted in the long term by better utilizing reserves of straight coking coal and increasing the life of the mine. The top seam unit was obtained on a rental basis and by the beginning of 1967, two panels, the first having a face length of 215 m, a panel length of 345 m, seam section of 1 219 mm, and the second the same face length, a panel length of 453 m and a seam section of 1 067 mm had been successfully longwalled. At this stage a best month of 24 602 metric tons with an average of 17 364 metric tons per full working month could be reported. Results were sufficiently encouraging for the company to exercise its right to acquire the equipment as its own asset and from 1967 onward, the unit has operated on this basis. From the point of view of rock mechanics, the dolerite sill over the first two panels was known to have 'bridged', only the material below the base of the sill having truly 'goafed'. No adverse pressure effects had significantly affected face operations except for two 'bumps' towards the end of operations in each panel. Gate road maintenance presented no problem. The strong inflows of water experienced on the first face were now known to be associated with an isolated water bearing fault plane and were not experienced on the second face. TOP SEAM EXPERIENCE - MAY, 1966, TO DATE Panel 2 (February, 1966, to November, 1966) The operations in panel 2 (see Fig. 1 for location of panels) continued uneventfully until the final month in this panel. At this stage, the chocks behaved most inconsistently often lowering under the weight of the canopies alone. Roof trouble became so severe that the unit was unable to produce adequately and blasting of chocks became a daily occurrence. Finally 24 m short of the planned limit it was decided that a complete overhaul was the only solution and the extraction operations commenced. At this time, the only fatal accident associated with longwalling occurred when a Non-White inadvertently fell into the panzer conveyor transfer point at the main gate. Some gate road difficulties in this panel were overcome by bolting tapes to the roof as decking and allowing them to pass over the gate chocks and collapse into the goaf. Two cribbed roof falls were also negotiated by the gate chocks at the face ends without undue difficulties. It was decided in the light of maintenance problems encountered on this face that the best solution would be to appoint a foreman solely in charge of face maintenance crews. This step has to a large extent contributed to success achieved subsequently. Panel 2-face length 213 m panel length 456 m cutting height 1 092 mm average monthly output 16 443 t best month 19 130 t Panel 3 (December, 1966, to June, 1967) By the end of 1966 results were sufficiently encouraging for the company to exercise its right to acquire the equipment. This was negotiated and the unit has operated as such to the present date. Prior to installation in panel 3, all face equipment was thoroughly overhauled underground and the chocks boost-tested to yield-load. The costs of overhauling, including almost complete hose replacement, were much higher than anticipated. Advice had been received that complete hose replacement was common practice after +/- 18 months' operation. Shortly after commencement of operations in panel 3 the chocks again failed as they had towards the end of panel 2. At this stage, the suppliers advised that the valve gear should have been repaired during the overhaul as the needle and seat seals could no longer be relied upon. This operation was impossible to carry out on the face and would have been e1ttremely time consuming. Fortunately a new type of valve gear was then available from the suppliers, termed the capsule-type valve, and a complete change over could be made on the face. The encapsulated valve has a delrin pad sealing over a raised

By J. A. L. Napier

An analytical solution to the best choice of cutoff grade for a mine is deduced for an assumed model of selective mining. It is found that this choice is determined by the constancy of a special function, the Hamiltonian, over the life of the mine. It is suggested that the appropriate constant value for this function should be determined by the expected long-term average of the paylimit. In that case, the optimum profile can be calculated at each point in time according to the prevailing working cost and product price. In particular, a decision can be made between operation at the true paylimit or at the subsidized paylimit implied by either the lease and tax formulas or the assistance formula for gold mines. It is also shown how the basic formulation can be extended for 'use in the determination of the optimum joint paylimit for two products over the life of the mine.

Jan 1, 1983

By R. M. Falcon

The paper opens with a summary of the major factors responsible for spontaneous combustion in coal discards. These include heat arising from the oxidation of coal and pyrite, and from the weting of dry or oxidized coal and pyrite; and certain petrographic conditions leading to self-ignition of the coal such as low rank, the presence of a high proportion of reactive constituents, and the occurence of highly porous and botanicaly structured inertinite macerals, The analyses of 6 samples from 2 discard dumps in the Witbank coalfield are included, which show that, according to a three-grade classification of liability to spontaneous combustion, one dump is 'reasonably liable' and the other is 'most liable', It is pointed out that factors other than liability to self-ignition determine the possible utilization of such discards, and that the prevention of oxidation is of great importance if they are to be utilized and spontaneous combustion to be prevented,

Jan 1, 1986

By T. van Heerden

A range of supported gold catalysts was prepared by ion exchange, varying many of the preparation variables, including concentration in the precursor solution, washing procedure, as well as drying and calcination procedures. These catalysts have been characterized extensively. TEM images show essentially the same crystallite size distributions, between 2?5 nm, for almost all catalysts, the only exception being catalysts not washed in ammonia, which did not show any small crystallites. Additional characterization with SEM yielded an interesting discovery. Catalysts that appear identical on the TEM also contain some large crystallites in the range of 50?500 nm. Differences in dispersion due to the drying procedure not seen on the TEM can now be observed. Oxygen chemisorption is being investigated as an additional method to characterize gold based catalysts to complement the typically used electron microscope techniques.

Jan 1, 2012

By T. J. Otto, C. H. Cloete, T. Mkhatshwa, T. J. van Heerden

The value realised by an open pit mine depends on the quality and integrity of the mine planning process as well as the level of execution against these mine plans. When managing the execution against the mine plan, how well the mine plan is executed spatially, is of critical importance. The short-term mine plan is the plan that is physically executed on most open pit mines. At the same time, the spatial mine-to-plan compliance is typically reconciled against the annual business plan. Short-term mine planning, therefore, plays an important role in the effectiveness of the spatial mine-to-plan compliance reconciliation process by providing operational teams with detailed designs and schedules, while ensuring that mining execution is spatially aligned with the business plan. The Sishen open pit iron ore mine (Sishen) strives to continuously improve spatial compliance to the business plan. Sishen enhanced the role that short-term mine planning plays in enabling the forward-looking component of the spatial mine-to-plan compliance reconciliation process. The enhanced short-term mine planning process focuses on detailed tactical sequence designs per mining pushback, the health of value chain buffers, spatial plan-to-plan reconciliation, and the associated management routines. This led to improved spatial mine-to-plan compliance to the business plan from 71% to 94% over the four-year period from 2020 to 2023 inclusive. These results indicate that the application of short-term mine planning as part of an integrated spatial mine-to-plan compliance process at open pit mines can contribute positively to improving the level of spatial execution against the business plan. This paper presents the enhancements made to the short-term mine planning process, which other open pit mining operations can consider to improve spatial mine-to-plan compliance.

May 27, 2025

A number of issues relating to the computational aspects of pillar design are addressed in this paper. The computation of average pillar stress values is important when attempting to establish criteria for pillar design and in the analysis of the stability of tabular pillar layouts. One of the default ?classic? numerical methods that are used to determine pillar stresses is the displacement discontinuity method. In many instances it is not clearly understood that this approach does suffer from some limitations, particularly in relation to the fact that in coarse element simulations, the simulated average pillar stress (APS) can depend on the chosen mesh size. The nature of this error is highlighted in this paper and some strategies are suggested to bound the magnitude of these errors. It is demonstrated as well that the popular linear stiffness approximation to pillar or seam compressibility does appear to allow reasonably accurate estimates of the average pillar stress when either the pillar height is varied or when the seam modulus differs from the host rock modulus. A practical implication of this study is that if the seam modulus is noticeably lower than that of the host rock, such as for coal seams, it is important to use a linear stiffness constitutive model for the pillars rather than a ?rigid? pillar assumption. This added complexity seems unnecessary, however, when simulating hard rock pillars in mines where the seam modulus is very similar to that of the surrounding rock.

Jan 1, 2011

By J. Bai, C. Zang, W. Yuan, L. Zhuang

"This study evaluates the efficiency of using mechanochemical reduction to assist the extraction of zinc from zinc ferrites by alkaline leaching. The transition of zinc ferrite into a metastable state after mechanochemical reduction with metallic iron contributes to the ready dissolution of zinc from the activated zinc ferrites in alkaline solution. Zinc ferrites were mechanochemically reduced under conditions of Fe:ZnFe2O4 mole ratio of 2:1, using 5 mm diameter stainless steel balls as the activation medium at a mass ratio of balls to raw materials of 25:1. Subsequent leaching in 6 mol/L NaOH solution at 90°C resulted in more than 70% Zn extraction. These results may be used to further develop a hydrometallurgical process for recovering zinc from zinc ferrites in alkaline solution.IntroductionZinc is an important nonferrous metal required for various applications in batteries, solder, dielectric materials, and piezoelectric materials. High-grade, easily concentrated ores are becoming scarcer: to meet world demands for zinc, it is necessary to develop economical and environmentally safe metallurgical technologies that use secondary materials such as leach residues of roasted zinc concentrates and electric arc furnace dusts. The zinc in the majority of these wastes occurs in the form of zinc ferrite (franklinite, ZnFe2O4), which is very stable and insoluble in most acidic and alkaline solutions (Langova and Matysek, 2009; Youcai and Stanforth, 2000).Several pyrometallurgical and hydrometallurgical processes (Jankovi et al., 2014; Langova and Matysek, 2010; Morcali et al., 2012; Shawabkeh, 2010) have been proposed for the treatment of zinc ferrites. Pyrometallurgical methods are costly owing to high energy consumption and their emissions of toxic gases are considered an environmental problem; hydrometallurgical methods are preferred for process economy and environmental reasons (Copur et al., 2004)."

Jan 1, 2016

By B. van Beek, J. H. Zietsman, A. Lexmond, N. J. Andrew

"In platinum group metals (PGMs) smelting, the slag-to-matte ratios are very high (approximately 4:1 to 9:1). The slag temperature therefore has a strong influence on matte temperature, which is a continual concern due to the low matte melting temperature and its resultant high superheat. Maintaining a stable slag temperature is therefore a critical safety and operational issue. This project aims to improve the understanding of the slag bath temperature drivers so that a more effective slag bath temperature control strategy can be implemented.With the aim of developing a multiphysics model of the furnace, extensive data analysis was performed to develop an in-depth understanding of the process before the start of modelling work. In this work, the effect of fluctuating furnace feed composition on slag temperature was examined. The results are based on equilibrium calculations using FactSage, implemented with ChemApp.A method was developed, using an equilibrium calculation at 25°C, to convert the available chemical assays to approximate mineralogical assays in order to determine the feed enthalpies. Actual 2013 furnace feed data was used and the average specific energy requirement to produce slag and matte at 1600°C was calculated to be 671 kWh/t concentrate. The furnace energy input was then kept constant, and it was found that the slag temperature varies significantly as a function of the feed composition. These results were compared to actual slag temperatures measured on site. It was found that close management of the furnace energy balance is important for maintaining a stable slag temperature."

Jan 1, 2014

By А. Klemyatov, T. Vologdina, U. Gordeeva, S. Zbinets, А. Kalugin, N. Rozhdestvenskiy

The plants integrated in the PhosAgro group of companies perform processes from scientific research to the project, from ore mining to shipment of fertilisers and feed additives to consumers. An extensive process chain maintains smooth interaction at all stages from research to production. As an example, the paper shows the revamp of two beneficiation plants (ANOF-2 and ANOF-3), requiring a comprehensive approach to task solving. The intention of the revamp was to increase the total yield of apatite and nepheline concentrate, taking into account the gradual decline in the quality of ore being processed and ensuring the yield stability of final products during re-equipment of the plants. At ANOF-3, the crushing and grinding cycle was modernised, the production of nepheline concentrate was implemented, and the capacity of apatite flotation, dehydration, and drying was increased. At ANOF-2 a new process chain was implemented for the production of apatite concentrate in workshops previously used for the production of nepheline concentrate. The main building was withdrawn from operation, where ore grinding, apatite flotation and apatite concentrate thickening had been carried out. The total increase in apatite concentrate yield at two beneficiation plants was 1.3 times. A further increase in concentrate production is planned. Keywords: Apatite, nepheline, fertilisers, beneficiation, flotation, column flotation machines, revamp, fine screening

Jan 1, 2020

By F. Smit

Mining and plant disciplines can benefit from pre-concentration of UG2 platinum ores in a dense medium separation (DMS) plant. The main benefit of DMS is the selective discard of waste from the feed stream, thus improving project economics while reducing the power and water requirements for processing. On the mining side, with the correct combination of variables, implementing DMS as a pre-concentration step can reduce cut-off grades and, as a result, increase ore reserves. In addition, the use of DMS technology gives the mine team the opportunity to consider alternative mining methods for an operation. The challenges of balancing the mining tonnage, methods, and resulting grades with the process plant size, recoveries, and DMS discard volumes have resulted in the development of a model to optimize value on new projects or existing operations. This paper will focus on these benefits from a mining and plant perspective, quantifying the water and power savings, and look at the economic benefits of pre-concentration and the effect of metal prices, plant feed grade, and DMS efficiency in different mining environments.

Jan 1, 2012

By A. J. Vietti

"Drought conditions and the possible lack of sufficient raw water supply are a constant reminder to the Southern African diamond miner of the need to minimize raw (new) water input while maintaining or expanding mine throughput.Since water input volume is directly related to water output (loss) volume, mainly to the tailings, it is obviously beneficial to dewater the tailings as much as possible. However, this may not always be as easily achieved as first thought. Owing to the unique mineralogy of each kimberlite deposit and the unique chemical profile of each raw water source, dewatering of the slurry is not a standard process that can be applied equally across the entire diamond industry.This paper presents a first-step strategy that kimberlite diamond mine operators should consider and apply to meet and exceed their watersaving targets. Introduction A common metric for measuring water use efficiency in a metallurgical plant is the amount of raw water used per ton of head feed treated (m3/t). The metric includes all water recycled from the dewatering process and from the tailings storage facility (TSF), and the only inputs are tons of ROM ore and cubic metres of raw water make-up to replace water lost to evaporation, seepage, and lock-up at the TSF (Figure 1).The volume of new raw water available to the process is not unlimited and a maximum monthly allowance is typically set during the mine permitting process. Consequently, if the volume of water lost exceeds the volume of new water available or alternatively, in times of drought where the allowable volume is unavailable, the plant has no other option but to reduce throughput.However, if the plant operator is able reduce the volume of water lost to the TSF, the plant will have additional capacity to either increase tonnage throughput or sustain nameplate capacity during periods of severe drought. The obvious conclusion, therefore, is to reduce the volume of water lost to the tailings by improving the efficiency of the dewatering process (typically gravity thickening) simply by increasing the density of the tailings being discharged. In practice, however, the unique mineralogy of each kimberlite deposit and the chemical profile of the raw water source used imply that at some mine sites, solid/liquid separation (or settling) of tailings slurries at the dewatering process is more easily achieved than at others. In these cases, simple practical steps such as firstly improving the flocculant reagent make-up and dosing control, and secondly focusing on thickener discharge control, are the only requirements for enhancing water recovery at the dewatering circuit. At other mine sites however, solid/liquid separation of the tailings may not be achieved at all, in which case efforts to increase water recovery by improving reagent dosage and thickener control would be fruitless unless the colloidal properties of the thickener feed slurry are modified."

Feb 1, 2019

By F. Reyes, M. Hilden, G. Forbes, M. Yahyaei

Since AlphaGo defeated the best human player in the game of Go in 2016 - something thought to be an impossible achievement - machine learning, deep learning and reinforcement learning have caught the attention of scientists, the media and the general public alike. This recent development in machine learning (let us remember that neural networks have been available for decades now) opens new possibilities in the areas of modelling and process control. However, little has been reported in the minerals processing industry about these new possibilities and opportunities. This paper presents and introduces the topic of reinforcement learning and the different platforms that allow its implementation. The paper shows the consequences of using reinforcement learning for the specific area of process control and optimisation. As an example, the authors present the use of reinforcement learning for the control of a SAG mill closed with a pebble crusher using industrial data from an operation. A comparison with advanced model-based process control is provided, showing the benefits and shortcomings of reinforcement learning as a new control and optimisation technique. Keywords: Reinforcement learning, deep learning, process control, process optimisation

Jan 1, 2020

Review Mining annual review June 1979. London, Mining Journal, 1979. 624 pp. Available as part of the subscription to the Mining Journal or at £12,50 per copy. This annual review is recognized internationally as the world leader in reporting on metals and minerals, and is eagerly awaited each year for its authoritative content, which is compiled by well-respected experts throughout the world.

Jan 1, 1979

By R. T. Muller

Contribution to the above paper by R. T. Muller (Visitor) Demands are being made and pressure is being brought to bear on employers in South Africa to narrow or eliminate the wage gap between their White and Black employees. Few thinking people will dispute the fact that a constant improvement in the wages of Black personnel is essential, but this cannot be done merely by a redistribution of income in South Africa.

Jan 7, 1974

By A. A. Bradley, P. J. D. Lloyd, A. J. Freemantle

Developments in centrifugal milling by A. A. BRADLEY,B.Sc. (Rhodes) (Visitor), A. J. FREEMANTLE, B.Sc. Hons. (Wits) (Visitor), and P. J. D. LLOYD, B.Sc. (Chem. Eng.), Ph.D. (Cape) (Fellow) The theoretical and experimental results of a study of centrifugalmills are reviewed. Means for the scaling up of such mills are suggested, and tests on a prototype mill are described. Despite the high lining wear in the centrifugal machine, which reduces its productive availability, its other advantages suggest that there is merit in pursuing its further development.

Jan 6, 1974

By M. J. Nicol, R. P. Colborn

An investigation into the kinetics and mechanism of the oxidation of iron (11) by oxygen in aqueous chloride solutions By R. P. COLBORN Pr. Eng., Ph.D. (Natal) and M. J. NICOLt B.Sc. Hons., Ph.D. (Rand) A reaction mechanism for the oxidation of ferrous chloride by molecular oxygen is developed. and the appropriate constants are evaluated. The effect of a number of cations and anions on the rate of oxidation has been tested.

By B. J. Madden

The formula developed by Salamon and Munro in 1967 has been used for the past 22 years in the design of bord-and-pillar workings in South African collieries. This paper re-assesses the formula In the light of Its past performance and examines the effect of mining method on coal-pillar strength. An account is given of the extension of the formula so that it can be applied beyond the empirical range for which it was derived, and the advantages of the extended formula (the squat-pillar formula) are described.

Jan 1, 1991

By C Smit

SYNOPSIS A prototype collimated radiometric face scanner was tested in the Harmony Gold Mine. The results obtained during the pilot study indicate that in situ radiometric uranium assays are statistically indistinguishable from those obtained conventionally from channel chip samples. In addition, the study demonstrated that reasonably reliable gold estimates can be deduced from the radiometric measurements, by use of the ratio of gold to uranium within a mine. The instrumentation, calibration procedures, and background determination are described briefly.

Jan 1, 1985