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By T. Tholana, P. N. Neingo

"A mineral resource constitutes the principal underlying asset of a mining company, and must be exploited and managed in such a way that maximum value is derived from it. Various asset management frameworks applicable to physical assets are available. This paper focuses on the extension of an asset management approach, particularly the Publicly Available Specifications (PAS) 55 asset management framework and the International Organization for Standardization (ISO) 55000 series of standards, to mineral assets. It is concluded that PAS 55 and the ISO 55000 series of standards can be extended to manage mineral assets, resulting in an integrated approach to sustainably optimize value from the company’s mineral assets. The benefits include improved returns on mineral assets, maximum mineral asset utilization, creation of an organizational culture focused on quality and continuous improvement, and assurance to stakeholders that the mineral assets are being efficiently managed over their entire life-cycle. In the context of this paper, a ‘mineral resource’ refers to both Resources and/or Reserves as classified by the SAMREC Code. IntroductionMining involves the extraction of valuable mineral resources from the Earth’s crust. A mineral resource is unique among natural resources in the sense that it is a wasting asset, which means that its value depreciates continuously with extraction. In addition, it is often impaired from time to time as commodity price changes (Macfarlane, 2011a). Since a mineral resource is a wasting asset there is a single opportunity available to optimize value from its extraction. Mine planning involves identifying a strategy to exploit the mineral resource in a way that maximizes value at an acceptable risk level throughout the life of mine (LOM). This optimum extraction strategy changes as economic conditions change, particularly commodity prices, as well as with improvements in orebody knowledge as mining progresses. It is the mineral deposit that competitively distinguishes two different mines and dictates the location of a mine; hence the adage ‘the orebody dictates’. Therefore, a mineral resource is the primary tangible asset possessed by mining companies and must be managed in such a manner that maximizes its value at an acceptable risk level (Macfarlane, 2011a)."

Jan 1, 2016

By J-C. Lv, S. Zhao, F-L. Zhang, T. Yan

"To eliminate nozzle clogging during continuous casting of ship plate steel and reduce the calcium addition required, the composition of CaO-MgOAl2O3 inclusions in the liquid phase region must be controlled. Thermodynamic calculations using FactSage show that when the balanced Al content is 0.03 mass %, aCa in the liquid steel is 1–10 ppm, and aMg is 0.01–0.3 ppm; and when the balanced Al content is 0.06 mass %, aCa in the liquid steel is 1–30 ppm, and aMg is 0.05–0.9 ppm. The CaO-Al2O3-CaS inclusions can be controlled in the liquid region below 1600°C at aCa values of 5 ppm, 15 ppm, and 30 ppm, and the balanced Al content should be controlled in the range of 0.001–0.050, 0.005–0.250, and 0.010–0.50 mass %, respectively. Based on theoretical analysis, the refining slag composition should be controlled to 45–53 mass % CaO, 30–37 mass % Al2O3, 6–12 mass % SiO2, and 6–8 mass % MgO. Optimization of the slag system and steel composition can reduce, or even eliminate, the requirement for calcium treatment. The total oxygen content in the slab can be controlled to 15–21 ppm, with typical micro-inclusions ??10 ??m in size, with compositions in the CaO-Al2O3-MgO and CaO-Al2O3-CaS systems. Introduction With the rapid growth in steel output in China and the changes in the macroeconomic environment, competition in the steel industry has become more intense. The technology for economical cleaning of high-grade steel is important for improving product quality and reducing production costs. However, the functions of refining equipment have not been fully taken into consideration, and the combination of refining methods has not been optimized completely. These problems have become a bottleneck, restricting modern steel enterprises from developing an economic clean steel production process. Ship plate steel produced by the converter long process has been selected as the example in this study. The composition of the steel is shown in Table I. This steel is a high-strength low-alloy steel for rolling media and heavy plates, mainly used for ocean-going vessels and other steel structures. The carbon content in the steel is far from the peritectic point, and the crack sensitivity of the slab surface is weak, so impurities and gas content are strictly limited. Therefore, the process route BOF (basic oxygen furnace) – LF (ladle furnace) – CC (continuous casting) needs to be studied."

Jan 1, 2018

By M. Berjak, D. M. Hawkins, V. E. Marting, R. B. MacGillivray

Contributions to discussion V. E. Marting (Associate Member): The authors of this paper should be congratulated in having pioneered the application of the Sharapov predictor to local data, and in having made a valuable contribution to South African experience in the quantitative treatment of geological information using a digital computer. Although the existing set of three programmes has been described as a 'suite' it does not represent an integrated system in the accepted sense since the communication between runs is made in the form of punched card files to enable the mining engineer to exercise discretion and control over the data entered at each phase. Although the authors are justifiably proud of the successful results obtained through the application of this method, I feel that it is my duty to comment that the computer aspect of the work has not reached full development and is deserving of the financial support needed to increase its widest usefulness. At the present moment a routine to detect 'missing information' has not been commissioned. Due to the lack of documentation the use of the programmes is restricted since the format for the submission of data is known only to the authors. It is very unlikely that a profitable return on the development costs will be enjoyed before the programme is fully documented and made available to wider use. My criticism of the present incomplete state of the work is being voiced purely with the intent of focussing on the significance of this project and the fact that it is deserving of development on an industry basis. It should be noted that at the economical cost of R4.00 per sq ft of computer print-out, a mining engineer can hardly afford to miss the opportunity of using the programmes on other (e.g. metalliferous) mineral deposits of tabular formation. Since the publication of this paper I have had an opportunity to see the results of the programmes developed by Drs M. D. G. Salamon and F. Deist for designing bord and pillar workings. The basic data for these latter programmes are extracted from iso-line plans drawn conventionally by geologists, a task that could be considerably expedited through the use of Mr Hawkins' programmes. It appears that the efforts of these respective teams are complementary and it is in the interests of the industry that these individual components should eventually form part of an integrated system. E. Margo* (Member): This paper is a good example of how and where the backroom scientist-in this case the mathematical statistician-can aid the mining engineer to develop not only more accurate means for dealing with valuation problems but to develop a tool that can also cope with changed geological specifications.

Jan 1, 1969

By W. Bleloch

Published in the Journal, September 1970, and presented at a colloquium on 'Future of the Steel Industry', 17th March, 1971. Errata (a) The final sentence in the penultimate paragraph, left hand column p. 34 of the text should read: 'In the hot regions of the bosh and stack the C+CO2=2CO reaction proceeds to CO from coke carbon and carbon dioxide generated by reduction of iron oxides by CO (the so-called solution loss of carbon)'. The italicised words have been omitted in the text. (b) Temperature in Fig. 3 is K not C. (c) In Fig. 9, connection from top plate of column 10 to heat exchanger 9 has been omitted. DISCUSSION J. B. BEEDON AND C. J. JONKER* Dr Bleloch's paper has outlined the advantages of recycling stack-gas of low nitrogen content in the blast furnace. In the first instance it would appear to be extremely beneficial in controlling the hearth temperature and allowing a 100 per cent oxygen-blast to be used. On a second point however we have to disagree. It is claimed that heat is liberated in the combustion zones by the oxidation of CO to CO2 near the tuyeres and in the subsequent reduction of the CO2 by CO. The author has hypothesized as follows, and we quote: 'The C+CO2=2 CO reaction in the outer combustion zone at about 1 700°C is no longer endothermic due to loss of stability of CO2 and its consequent dissociation into CO and free oxygen.' This statement requires closer examination. If we consider the individual events occurring with the CO in the recycled topgas, with the oxygen in the blast and with the coke in the burden, as done by Dr Bleloch, the following heat considerations emerge. All heats of reaction are calculated from the enthalpy values of each reactant and product, as well as the temperature dependence of their molar heat capacities. (a) At the tuyeres: CO+1/2 O2-+CO2 . . . . . (1) CO, CO2 at 500°C; O2 at 25°C. ?H reaction = -65 000 cals/mole CO The reaction (1) is exothermic. (b) In the outer combustion zone: CO2+C?2 CO . . . . . . . . (2) C, CO at 1 700°C; CO2 at 500 0C. ?H reaction = -56 000 cals/mole CO Reaction (2) is endothermic. The overall heat gain of reactions (1) and (2) is ?H=+56 000 - 65 000= -9000 cals/mole CO i.e. the overall reaction is slightly exothermic, the amount of heat liberated is however smaller than suggested by Dr Bleloch. (c) The dissociation CO2+CO+1/2 O2 . . . . . (3) is a highly endothermic reaction at 1 700 0C, requiring 82 000 cals/mole CO2 to proceed to completion. Furthermore, this reaction (3) is energetically not favoured to proceed, as can be observed from the free energy-temperature relationship of figure (3), which favours the formation of CO2 at 1 700 0C. Hence direct dissociation of CO2 at 1 700 0C appears to be unlikely.* The concept of recycling topgas is however still intriguing from many other points of view. Apart from providing a very efficient method of hearth temperature control, it will create a much higher CO partial pressure in the stack. Whether the rate of indirect reduction in the stack will be proportional to this partial pressure or not is un-

Jan 9, 1970

Solvent loading: It is agreed that it is advantageous, as far as reagent cost is concerned, to operate at maximum solvent loading. However, owing to the fact that no reliable automatic solvent flow controller is yet available, control tends to be erratic with the result that there is risk of losing uranium. As can be seen from Mr Tunley's plot, the graph representing total ammonia cost is fairly flat between loadings of 3.5 and 4.0 gpl, while the graph representing uranium loss in the raffinate is steep. It is, therefore, safer to operate slightly below optimum loading in the absence of absolute control. Should the purity of the final product gain importance, operation at optimum loading or even slightly above, may be advantageous. Solvent losses and clarification of pregnant solution: The cost of replacing solvent, as indicated by Mr Carman, represents a high percentage of the overall Purlex costs. It is, therefore, agreed that, amongst other factors, efficient clarification of the pregnant solution and mixer-settler design have an important bearing on the cost. However, as at Blyvooruitzicht, the Buffelsfontein plant can tolerate a suspended solids content of 50 ppm without serious deterioration in efficiency. An increase in the suspended solids content from 15 to 50 ppm normally accounts for an increase of approximately 1 ppm in amine loss. Clarification of acidic pregnant solutions is much more difficult than that of alkaline solutions. Sand clarifiers are being used at Buffelsfontein with moderate success. Clarified solutions containing less than 20 ppm of solids in suspension can be produced, depending on the quantity and type of flocculant used in the filtration process, the thickness and the grading analysis of the sand beds and the quality of supervision. An on-stream duty of approximately 0.15 gallon per sq ft of surface area is obtained. Hydraulic handling of clarifier sand, while cleaning, may effect the efficiency of a clarifier, as size segregation of sand particles may lead to inefficient "pockets" in a sand bed. Control of the solution level in the sand clarifier also affects the quality of the effluent. Exposure of the sand bed invariably causes a cloudy effluent. Recycling of the final and initial solution discharged before and after skimming the clarifier sand improves the clarification efficiency significantly. Post-precipitation of silica and possibly sulphate salts, after clarification, may give a false impression of inefficient clarification. This phenomenon is more pronounced when treating hot supersaturated solutions. Ageing and cooling of the solution before clarification assist in stabilising the solution and reduce precipitation after clarification. The maintenance cost of sand clarifiers at Buffelsfontein is in the order of 0.35 cent per metric ton of solution treated. A redundant ion-exchange column has been tested successfully, either as a primary or a secondary clarifier, depending on the quality of the filtrate treated. The conventional pebble and sand beds in the column have been retained. Operation was enhanced by the addition of a two-foot deep resin bed as a filtering medium. The suspended solids content was reduced to less than 5 ppm. Mixer-settlers: Mixer-settler design is a controversial subject but, in view of the high costs of solvent compared to relatively low pumping costs, it is believed that mixing and pumping functions should be separated. Separate facilities for pumping and mixing not only give greater operational flexibility but also permit eventual changes in the flow circuit as technology develops. Scrub technique: Unlike the pilot plant, where the crud was carried away in the aqueous phase, crud is carried over with the organic phase in the full-scale plant. This carry-over of crud into the strip bank aggravates crud formation in the first strip stage. This phenomenon may result from a slight difference in settler design and/or a variation in the relative phase depths in the two plants. Bypassing of the scrub aqueous solution from the fourth to the second stage was introduced in order to avoid dilution and partial neutralisation of the acid added to the third stage, as it is known that a low pH value enhances the transfer of iron into the aqueous phase. The low aqueous flowrate in the third scrubbing stage was overcome by introducing an aqueous phase recycle. Gleaning: Mr MacDonald's contribution on "gleaning" was most informative. This process for recovering solvent from raffinate seems to show at least as much promise as coalescers and other methods employed to date. Similar tests conducted at Buffelsfontein pilot plant gave 60 per cent amine recovery from raffinate containing 10 ppm amine. However, the circuit was not operated at equilibrium and recovery may be expected to drop as the amine concentration of the "Gleaner" circuit increases.

By Ronald J. Roman

All mining and milling complexes can, within limits, operate at a variety of production rates. Factors of economic importance that are affected by production rates are costs per unit production, recovery and product grade. Once the relationship between the production rate and cost is established, dynamic programming techniques can be used to define the optimum production schedule (that is, the production schedule that maximizes the present worth of the operation) for the life of the deposit. In general, assuming that the capacity and efficiency of the physical plant remain unchanged throughout the life of the deposit, the optimum production rate is not constant for the life of the deposit, but declines gradually as the deposit is being exhausted.

Jan 1, 2014

Determination of mineral composition is of significant importance for the mineral sands industry as this information is required for orebody evaluation, the control of mineral processing plants and the determination of product quality. Mineral processing is made more difficult by the presence of pseudorutile, leucoxene, altered-ilmenite and particle inclusions. The paper investigates the properties of such elements from a single product stream using an MLA approach.

Jan 1, 2007

By P. Smit

Health hazards of South African . mIne water by P. SMIT, M.B., Ch.B., D.T.M.H., D.P.H. (Visitor) The water supply in South African mines is not as good as most mining engineers would like to believe, nor is sufficient care taken to ensure the safe dispensing, transportation, and consumption of water. As long as cases of dysentery and typhoid continue to occur in large numbers, it can be assumed only that the water supplies are not safe, or that the water-dispensing methods are poor. A dual water supply is a real health hazard. The ideal water supply is a single safe source for human consumption and industrial use.

Jan 4, 1977

By A. L. Hinde, P. J. D. Lloyd, D. Hebden

The philosophy behind the design of a rougher cell for the handling of coarse feed and for the maximization of recovery within a given time is presented. The application of this philosophy in cells of 22-litre, 320-litre, and 3 m3 capacity is outlined, and the results of the process testing of these cells are given. In this testing, particular accent was placed on the recovery of water as a guide to the recovery of the finest size of particles, and the rate of recovery of various mineral species. It is concluded that a promising concept has been identified, but that further testing is needed for the firmer establishment of optimal operational parameters in a production environment.

Jan 1, 1983

By C. J. Alberts

The perception exists in many organizations that, once an energy management system is employed, there need not be any further adjustments to such a system. This is not true, as the field of energy management is a dynamic field, and constant feedback and adjustments are needed to enhance the efficiency of a system to the point that substantial pay-back can be obtained. It is stressed that for any energy management system to be successful, careful planning should permeate through the organization, from the top level decision-makers to the workplace. This paper describes a dynamic technique by which energy management strategies can be evaluated by viewing the effects of the strategies on the organization as a whole.

Jan 1, 1997

By K. R. Johnson

Changes in the pattern of computer systems for mine planning and geologic evaluation are occurring as a consequence of the introduction of 3-D high performance graphics terminals. The VULCAN system is described and an outline of its functionality is discussed. Increasingly, the integration of both database systems and graphics and databases will occur, allowing a greater degree of information access to mine operators. The system covers areas of database, drillhole graphics, mine design and layout, surface and solid modelling as well as reserve inventory generation. The design of the software is such that it is fairly commodity-independent and entirely menu-driven, to enable non computer-trained personnel easily to access the system.

Jan 1, 1987

One of the most critical and complicated steps in mine design is a suitable mining method selection based upon geological, geotechnical, geographical and economical parameters. Since there are many factors involved in mining method selection, the decision-making process is very difficult. In this paper, the Analytical Hierarchy Process, with 13 criteria, is used to develop a suitable mining method for the Golbini No. 8 deposit in Jajarm (Iran). Six alternatives (conventional cut and fill, mechanized cut and fill, shrinkage stoping, sublevel stoping, bench mining, and stull stoping) are evaluated. The studies show that the suitable mining method for this deposit in the present situation is the conventional cut and fill method.

Jan 1, 2008

Written contribution by J.F. Curtist The paper is an interesting exercise in the application of mathematics to a theoretical situation presumed to oc¬cur underground. However, some of the assumptions on which the calculations are based are unlikely to occur in any actual mining situation. Assumptions made in the paper Objections to these assumptions are noted seriatim. (1) `it is assumed that any loading-rate effect will be small and can be ignored' (p. 3). A loading rate of the order of 50 kN-12 per cent of the maximum load-is customary to prevent the props being dislodged by the blast. (2) `it is unlikely that all the props will be installed with their full stroke available, but for the purpose of this analysis, it is assumed that this is the case' (p. 3). A more reasonable assumption would be that, with the most careful selection of prop length and the most rigid control of stope width, the available stroke length would on average be reduced by the order of 70 mm on installation-20 per cent of full stroke. (3) `each support unit supports a block of 3,3 m2 bound¬ed by vertical fractures with zero cohesion that are parallel and at right-angles to the stope face respec¬tively' (p. 5). Joughin and Jager have shown this to be an unwar¬ranted assumption in respect of the positioning, the direction, and the dip of the fractures present in a stope in rockburst conditions. If, as assumed, there was `zero cohesion' between such blocks, the removal of a hydraulic prop in row 3 supporting such a block would ensure that it fell out before a profile prop could be installed in the same position. (4) `the height of each of these blocks is 3,0 m, the volume of each block is 9,9M3, the mass of each block is 27 225 kg' (p. 5). The concept of individual props being affected by dis¬crete blocks of known and uniform dimensions of relatively low ground velocities (less than or equal to 2,16 m/s) acting in line with the axis of the prop is one that lends itself to neat mathematical calcula¬tions. It does not, however, bear much similarity to the actual conditions that are exposed underground in excavations where rockbursts occur. Ortlepp and co-workers concluded that `the impulse of the shockwave can cause the displacement of the

Jan 1, 1991

By C. Musingwini, T. C. Sebutsoe

"The primary objective of any mining business unit is to make profit by extracting, processing, and selling minerals from a particular mineral deposit. It is important to optimize extraction of the mineral resource given time, capital, and space constraints. Mineral extraction is often associated with uncertainty due to variable technical and human factors. Technical factors such as grade distribution, ground conditions, and equipment reliability influence the performance of a mining production system (MPS). The performance of the MPS is also impacted by human factors such as employee skills, health, and attendance. Uncertainty associated with technical and human factors often leads to actual output differing from what was planned. Therefore an in-depth analysis of the significant causes of deviations from the planned outcomes is relevant.In this paper we investigate the empirical relationships between inputs and outputs in a MPS in order to assist management in directing efforts at key production drivers. Once a MPS has been empirically characterized, more effort and resources can be focused on the key decision-making variables (DMVs) in order to meet the planned outcomes.For a typical underground, tabular, narrow-reef conventional drill-andblast platinum mine the face advance, face length mined, number of teams, team efficiencies, and team size have a statistically significant relationship with the centares (m²) produced, which is a key performance indicator (KPI). The production function derived in this study can be used to align the physical, technical, and human factors together to predict the optimal output level. The most significant production lever of the MPS is the face advance. IntroductionA mining production system (MPS) is a result of an iterative process of design, planning, and optimization of mining input variables and decision-making variables (DMVs). The MPS exists within the mineral extraction link of the complete mining value chain. It represents the stage where mining companies have the opportunity, through production, to start generating returns on the investments undertaken. Returns on shareholders’ investments can be realized at this stage."

Jan 1, 2017

On 8th August, 1973, 57 members of the Institute visited Prieska Copper Mine at Copperton in the North Western Cape. After tea, the group was welcomed by the Management of the Mine, and Mr Grobler, the General Manager, gave an introductory talk. After they had been taken on a bus tour of the property, the members split up into groups and proceeded underground or to the concentrator. Of special interest to the mining members were the rock-handling methods in use at the mine. The stoping method that has been adopted is the continuous, long-hole sub-level open stoping system. The broken ore is moved with rubbertyred front-end loaders from the stope drawpoints to the ore passes. On the main tramming levels, the ore is collected by electric locomotives and discharged into the main shaft system at the bottom. At the shaft ore-pass system, a gyratory crusher that can accept rocks up to 1 m in diameter has been installed. The crushed product (minus 150 mm in diameter) will then be hoisted by a Koepe winder, which has a rated capacity of 900 tons per hour. The mine is also served by a declined roadway from surface, which facilitates the movement of heavy equipment. The most striking aspect is the productivity that can be achieved by the use of a trackless-mining system. For the production of 230 000 tons per month, it is estimated that 6 stopers and 7 developers will be required. The total labour force will be approximately 2500-2000 Bantu, 200 Coloured, and 300 Europeans. Training of maintenance and operating personnel is of the utmost importance, and it was interesting to note the extent to which non-White labour is being used in these occupations. Lunch was served at the single quarters, after which the group proceeded to the airfield for the return trip to Johannesburg. On behalf of the visitors, Professor de V. Lambrechts thanked Mr Grobler and his staff for the most interesting morning and for their hospitality,

Jan 10, 1973

By J. Githiria, C. Musingwini, J. Muriuki

"The maximization of net present value (NPV) is a primary objective in open pit mine planning processes. In an attempt to meet this objective, cut-off grades are considered in all the stages of mining. There are three stages involved in resource extraction, namely mining, processing, and refining/marketing, which are all defined in Lane’s approach. Using Lane’s approach, the economics involved in each stage are identified independently and interact to provide an optimum cut-off grade.A hypothetical block model is used to illustrate how a computer application (Cut-off Grade Optimiser) that was developed in this study is used in the optimization of cut-off grades using Lane’s algorithm. This paper analyses the application of Lane’s approach for a single element in cut-off grade optimization as applied in the calculation of the optimum cut-off grade through a computer-aided application.Although Lane’s approach is not complex it is not widely applied to maximize the NPV of mining operations as the iterative calculations can be lengthy. The computer application developed in this study shows how Lane’s algorithm combined with a linear programming function can optimize cut-off grade with regard to the complex situations faced by mining operations. The application saves users from performing otherwise lengthy, iterative calculations.Background and previous researchAn open pit mining complex constitutes multiple pits, ore processing streams, and metal refining options. An open pit mine produces ore (valuable) and waste material. Ore is transported to an appropriate processing stream for producing concentrate. A refinery processes the concentrate to produce marketable metal product (Lane, 1964, 1988).Cut-off grade optimization is based on Lane’s theory, where a mining system consists of three major stages such as mining, milling, and refining. In this approach the cut-off grade selection is done between six cut-off grades that involve all stages of a mining operation. Selection of the optimum cut-off grade influences the net present value (NPV) by maximizing the net cash flows from the operation."

Jan 1, 2016

"Pressure leaching is an option for copper recovery from chalcopyrite. Leaching takes place at high temperatures and pressures in the presence of an oxidizing agent. This work shows how the performance of pressure leaching circuits may be improved by optimizing the configuration of autoclaves and the heat removal method. The objective was achieved by creating an autoclave model that combines mass balances, energy balances, and population balances. A base case circuit consisting of a single autoclave was chosen, and this base case was compared with options for increasing capacity by adding more autoclaves to the circuit. These capacity increase options include circuits in which the additional autoclave is added in parallel, in series, and in series with thickening between the two autoclaves. The copper extraction, productivity, and cooling requirements for these options are compared. The series circuit with interstage thickening has the highest extraction and productivity; however, the cooling duty in the first autoclave is high. In addition to comparing circuit options, three options for heat removal were investigated: cooling coils, quench water, and flash recycling. The flash recycle option results in the highest copper extraction, and the quench water option the lowest copper extraction. IntroductionBackgroundCopper sulphide ores and concentrates containing chalcopyrite are typically hard to leach under atmospheric conditions. Smelting of these ores may also be a challenge, due to the inability of smelting processes to handle low-grade ores, as well as impurities such as arsenic and antimony. Alternative hydrometallurgical routes to treating chalcopyrite are therefore required. One option is pressure leaching, which is the focus of this work.In the pressure leaching process, sulphide ores and concentrates are leached in autoclaves at temperatures of 200 to 220°C and a pressure of about 3000 kPa (Schlesinger et al., 2011). An oxidizing agent, usually oxygen, is required to oxidize sulphide ions to sulphate ions. The oxidation of sulphur to sulphate is exothermic; consequently, a method of heat removal is necessary in order to maintain the temperature in the autoclave."

Jan 1, 2016

By P. Ramphisa

In South Africa reprocessing used oil into industrial fuel is commercially attractive. The reprocessed product is sold as a substitute for heavy fuel oil that is derived from crude oil (Rose Foundation, 2008). The objective of this investigation is to evaluate the use of petroleum products such as oil, petrol and diesel in Driefontein Gold Mine. The recovery of oil through recycling has costs implications to the mine while the release of petroleum hydrocarbons due to the poor management of the use of petroleum products has potential to cause pollution and negatively affect soil and water. Environmental legislation such as the Minerals and Petroleum Resources Development Act (Act 28, 2002) requires that such negative impacts to the environment be rehabilitated. The rehabilitation cost of pollution caused by petroleum products to the environment increases operational costs in business. Field assessment is conducted to identify the main sources of petroleum product contamination such as leaks in piping, overfills and petroleum spills. The areas that require mitigation would be mapped in the study. Recommendations for mitigation measures will focus on using the best practicable environmental option (BPEO) as well as the best available technology not exceeding excessive cost (BATNEEC). The recovery of oil from the mining industry is generally low and very poor. In this study, the effect of proper management of petroleum products to the mine?s operational costs is investigated and compared to there mediation costs of the impacts caused due to poor management of these products. The study investigates the financial benefit that could be obtained from proper environmental practices and efficient usage of petroleum products in the mining.

Jan 1, 2009

By R. G. B. Pickering

Conventional mining is an excellent match of mining method and equipment. The result is guaranteed performance. Change in the form of different technology requires change in mining method until the best match is achieved. This paper covers the development of technology and mining method that has occurred in the last five years to arrive at the objective of successful mechanized mining in narrow reef platinum deposits of the Bushveld Complex.

Jan 1, 2004

By J. H. Lazarus

The evolution of backfill-distribution systems for deep gold mines is examined. The status of current technology for the design of backfill-distribution systems is considered with reference to pipeline pressure gradients, pipeline wear rates, and slurry flow regimes. The effect of an increase in the relative density of the slurry on the performance of backfill-distribution systems is illustrated. Future trends regarding the design of backfill-distribution systems are discussed.

Jan 1, 1993