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By M. Mabuza

Two Rivers Platinum Mine (TRP) is a 55:45 joint venture between African Rainbow Minerals and Impala Platinum, respectively. TRP is situated on the farm Dwarsrivier 372KT, in the southernsector of the eastern limb of the Bushveld Complex. Dwarsrivier is underlain by rock formations of the Critical Zone (Winterveld norite-anorthosite) and the Main Zone (Winnarshoek norite-anorthosite). Two economically significant PGE-bearing layers namely the Merensky Reef and the UG2chromitite are found within the upper Critical Zone. These are approximately 140 m vertically apart, separated by a series of barren norites and anorthosites. At Two Rivers Platinum, only theUG2 chromitite layer is currently being mined for PGEs. The UG2 sub-outcrops north/south along the Klein Dwarsrivier river valley with a very gentle dip to the west of 7º?10º. At Two Rivers the UG2 averages about 120 cm in thickness, overlain by a waste pyroxenite parting about 20 cm thick, above which is a leader seam about 20 cm in thickness. The mineable cut consists of the leader seam, the waste pyroxenite parting, the UG2chromitite layer including 25 cm of pegmatoidal pyroxenite footwall, below the UG2 reef to a minimum of 185 cm trackless mining cut. The mining method is a mechanized bord and pillar layout, which will yield a total of 185 000tons at full run of mine production from the underground UG2 mine.

Jan 1, 2006

By C. A. Cousins

Platinoids in the Witwatersrand system C. A. Cousins. (visitor) Study of the composition, nature and mode of occurrence of platinoid minerals in the conglomerates of the Witwatersrand System leads to some interesting conclusions. Genetic evidence strongly favours a placer origin, while comparison with modern alluvial platinoids supports this view but also indicates that they are very mature alluvials. The compositional patterns and distribution of the platinoids in the Witwatersrand basin suggest the water was saline. Study of the mineralogy was greatly facilitated by the electron-probe.

The auriferous Carbon Leader reef which is the subject of this study can be subdivided into two types of sedimentological units, each with its characteristic spatial variability. For the longwall mining method applied at Western Deep Levels gold mine, subdivision of the orebody improves the estimation of the actual error variance, but not the grade estimates themselves. This can only be achieved by choosing oreblocks of a size and shape that are maximal parallel to the longwall and minimal at right angles to it, within the limits of mining restrictions. While simple kriging is the only method that gives unbiased results, the difference between untransformed and logtransformed kriging is negligible.

Jan 1, 1987

Two column flotation cells have been installed as single-stage rougher cleaners at Batu Hijau. The objective was to provide additional residence time in the cleaner circuit to maintain maximum recovery during high mill throughput and high feed grade periods. Installation of the columns has allowed the relatively coarse, well-liberated rougher concentrate and the relatively fine, low-grade scavenger concentrate to be treated separately, improving metallurgical performance in the cleaner circuit. Plant results show that the columns produce a high-grade concentrate at an average stage copper recovery of approximately 80 per cent. There is also indication of increases in both overall plant copper recovery and final concentrate copper grade. In addition, the operability and stability of the cleaner circuit has greatly improved.

Jan 1, 2007

By D. Rankin, H. J. Stucke, W. K. B. Loftus

SYNOPSIS Planning for the Finsch opencast mine is described with particular reference to the proposed stages of mining, the method used to determine the limits of surface waste rock stripping and the derivation of the waste rock mining rate. Actual mining practice is described and the possibility of installing a skipway into the open mine at a later date is discussed. In addition, the waste dumping policy is referred to: this includes the cutting of slots through the hill on the west side of the pipe to increase tramming speeds. The crushing, washing and recovery sections of the treatment plant are described together with the various changes being made to the circuit. The most important of these are the installation of X-ray sorters for the final recovery of + 7 mesh diamonds, and the vibrating grease belt for the recovery of - 7 mesh diamonds. INTRODUCTION The Finsch Mine, a subsidiary of De Beers Consolidated Mines, Limited, is situated approximately 90 miles west of Kimberley, and some two miles from the village of Lime Acres, in the Northern Cape. Lime Acres is a company village shared with the Northern Lime Company which operates a limestone quarry and works near the township. The mine is near the southern end of a well mineralized belt producing limestone, asbestos, diamonds, manganese and iron ore, and the whole area has developed rapidly in the last 20 years. Power for mining operations is drawn from the Electricity Supply Commission. The mine is conveniently close to the main Kimberley-Postmasburg railway line. Water is drawn from the Vaal River through a 57 mile State-owned pipeline at present temporarily operated by mine staff. DISCOVERY AND PROSPECTING OPERATIONS The diamondiferous orebody is situated on the farm Brits, which is State-owned ground that has been leased to the Company for the period of mining operations. There are indications of interest in the area by prospectors (Richter, de Bruin and others) over a period of some 30 years and the discovery would possibly have been made years earlier but for a law preventing prospecting for precious stones on State-owned land. Although the pipe formed a clear depression on the hill where it was situated, it was overlain by some 5 to 40 ft of banded ironstone rubble which was washed into the depression from the hillside above the pipe and obscured the deposit. In 1958, Mr Allister Thornton Fincham applied for, and was granted, the right to prospect for base minerals on the farm Brits. Asbestos is known to occur in the area and, initially, was Mr Fincham's primary interest. However, during the prospecting operations, garnets were discovered and it was realised that diamonds might be present. In 1960, when the Precious Stones Amendment Act was passed, a company

By F. J. J. Blaauw

This paper discusses a bold venture into the use of interactive computer graphics technology as the foundation of an overall integrated mining managerial system. The wide range of applications that can be addressed through the use of this technology is highlighted, and some discussion of the major facts of this project, which is still in its early stages, is provided.

Jan 1, 1987

The optimal design of production phases and ultimate pit limit for an open pit mining operation may be generated using conventional or stochastic approaches. Unlike the conventional approach, the stochastic framework accounts for expected variability and uncertainty in metal content by considering a set of equally probable realizations (models) of the orebody. This paper evaluates the performance of a new stochastic network flow approach for the development of optimal phase design and ultimate pit limit using a gold deposit as the case study. The stochastic and conventional frameworks as considered here utilize the maximum flow and Lerchs-Grossman (LG) algorithms, respectively. The LG algorithm is restricted to considering an estimated (average-type) orebody model, while the stochastic maximum flow algorithm is developed to simultaneously use a set of simulated orebody realizations as an input. The case study demonstrates that, when compared to the conventional LG algorithm as used in the industry, the stochastic approach generates a 30 per cent increase in discounted cash flow, a 21 per cent larger ultimate pit limit, and about 7 per cent more metal, while it maintains a consistency in phase size.

Jan 1, 2012

The block caving mining method is used to mine large orebodies in weak rock types. The layout for block caving requires that a considerable amount of development is carried out on the extraction drives and drawpoints at the bottom of the caving block. The development further weakens the rock mass, and stability problems arise in the extraction drives and extraction pillars. This paper describes the application of three-dimensional boundary element models in analysing alternative extraction level layouts for a block caving mine. The objectives were to establish whether the use of lateral pillars between drawpoints was warranted, and whether the lateral pillars should be offset or not. The stability of the rock mass within the pillars as well as the stability of the extraction drives were taken into account. The results showed that after undercutting, the horizontal and vertical stresses in the extraction level would be reduced. The lateral pillars would significantly improve the stability of the extraction pillars and extraction drives. There would be a minor improvement in stability if the lateral pillars were offset.

Jan 1, 1987

Traditionally, the term `total system' design applied to comminution conjures images of a design that addresses the usual business and operational drivers, including health, safety and environment (HSE), capital cost, operating cost, throughput and product size. In terms of the physical process, the mine to customer type studies also exemplifies an approach that satisfies a normal `total systems' approach. As an industry there are drivers that are forcing a re-consideration of what we regard as our total system. This modification of drivers will require changes in approach and implementation to provide the best solutions. Some of the major changes that impact the drivers are shortage of skilled personnel, changing social attitudes to remote mine site work, greenhouse gas emissions (GHGE), quality of available orebodies and geographical location of new mine sites. One approach to the challenges presented by the changing business environment is to expand the system envelope around comminution to ensure that comminution, in its widest sense, helps to meet the new realities. Background to the new drivers and constraints is provided and some of the appropriate responses are examined in relation to the impact on the total system.

Jan 1, 2007

By T. Oberthür

The Great Dyke of Zimbabwe constitutes the world&apos;s second largest reserve of platinum group elements (PGE) after the Bushveld Complex in neighbouring South Africa. Within the Great Dyke, economic concentrations of PGE are restricted to sulphide disseminations of the Main Sulphide Zone (MSZ), which are currently mined at the Ngezi, Unki, and Mimosa mines. Near-surface oxidized MSZ ores have a large potential. Their total resources are in the range of 160-250 Mt; however, all previous attempts to extract the PGE from this ore type have proved uneconomic due to low PGE recoveries (<< 50 per cent) achieved by conventional metallurgical methods. Within the ores of pristine, sulphide-bearing MSZ, the PGE are bimodally distributed. Platinum occurs mainly in the form of discrete platinum group mineral (PGM) grains (mainly bismuthotellurides, sulphides, and arsenides), whereas approximately 80 per cent of the Pd (and some Rh) is hosted in pentlandite. Within the oxidized MSZ ores, the PGE are polymodally distributed. Whereas the arsenide- and sulphide-PGMs that make up approximately 25 per cent of the original Pt content of the ore largely remain stable (relict PGMs), the remaining PGMs are disintegrated. The base metal sulphides are destroyed, partly releasing their base metal and PGE contents, and are replaced by iron oxides or hydroxides. Unspecified amounts of the PGE are redistributed and either form secondary PGMs, are found in chemically and mineralogically ill-defined (Pt/Pd) -oxides or hydroxides, or in iron-hydroxides, Mn-Co-hydroxides, and in secondary silicates. The problematic processing of oxidized MSZ ores is attributable to their complex nature and polymodal distribution of the PGE, prohibiting a significant upgrading of the ores by conventional metallurgical methods. Therefore, only bulk leaching methods are viable for ore treatment, and novel metallurgical methods have to be developed for the processing of these ores. Our ongoing work aims at locating the PGE in their mineralogical form in order to understand the mineralogical balance of the PGE in the ores and thereby facilitate the evaluation of metallurgical options for their recovery. A short overview on options and recent advances regarding the recovery of the PGE from oxidized ores is given.

Jan 1, 2013

By A. C. Legg

BCL Limited operates a nickel-copper flash smelting operation at Selebi-Phikwe, situated in north-eastern Botswana. The original furnace was commissioned in 1973 and produces a high grade sulphide matte, from its own concentrates plus additional custom feeds, which is shipped to refineries in Zimbabwe and Norway for further processing. In 2004 the company embarked on a study to modernize the flash smelting complex, with an emphasis on extending campaign life, improving the furnace integrity and enhancing throughput and process efficiency. First phase initiatives from this exercise, including the fundamental switch to a single concentrate burner operation of the flash smelting furnace, were implemented in July 2007. This paper summarizes experiences prior to implementation, details the specific improvements made to the flash smelting furnace and associated operations, highlights the improved plant performance after modernization, and outlines second phase expansion plans that are currently under consideration.

Jan 1, 2009

By G. P. Dyke

The MRMR classification system was developed specifically for mining applications, namely caving operations, and is one of three rock mass classification systems used in the South African mining industry today. Increased usage of the MRMR classification system has raised concerns that it does not adequately address the role played by discontinuities, veins and cemented joints in a jointed rock mass. To address these concerns, Laubscher and Jakubec introduced the In-Situ Rock Mass Classification System (IRMR) in the year 2000. Although the IRMR system is more applicable to a jointed rock mass than the MRMR system, a quantitative comparison of the MRMR and IRMR classification systems indicates that there is not a significant difference between the resultant rock mass rating values derived from the two classification systems.

Jan 1, 2008

By M. Gershon

During the exploration phase of mine development, drilling location is in the hands of the geologist, and drilling is spread over the entire domain of the property or deposit boundary. During the mine planning and production phases, however, the mining engineer also needs to use the information obtained from the drilling program. Thus it is important that he/she have some input into the selection of the drillhole locations. Usually, when exploitation will center on one portion of the property for the immediate future, the mine engineer will want to have more of the drilling performed in this area. In the early stages, the purpose of the drilling was to make a decision concerning whether or not to develop a mine. Once the decision is made to proceed with this development, the mine engineer needs to know, with as much confidence as possible, the estimate of the ore grade at each location where mining will occur. This paper addresses the problem of drillhole location for the needs of the mining engineer. The goal is to compare a variety of approaches to drillhole location with a new approach developed by the author. This approach is based on the theory of geostatistics, with the estimation variance used as the quantitative measure of the confidence attained. The goal of the algorithm is to select those locations that minimize the estimation variance subject to constraints on the number of drillholes and the total cost of the drilling program. The kriging equations are used as another set of constraints. The resulting nonlinear optimization model is expressed in terms of binary (0 - 1) decision variables leading to a solution by the branch-and-bound procedure. The algorithm has been applied at a number of operating properties, and practical results and suggestions for implementation are provided.

Jan 1, 1987

By L. J. Park, A. D. Thomas

Gladstone Pacific Nickel Limited (GPNL) is proposing to develop the Marlborough nickel deposit in central Queensland and the Yarwun Refinery in Gladstone, Australia. The project proposes to connect the two sites with a DN 500, single pump station slurry pipeline, transporting up to 4 Mtpa of nickel laterite ore as slurry 180 km from Marlborough to Gladstone. Slurry viscosity is a key parameter influencing pipeline design. Not only is the slurry viscosity relatively high, requiring pumping at low concentrations, around 30% by weight, but the viscosity (yield stress and plastic viscosity) increases with shearing or agitation. It is argued that the viscosity changes can be analysed in terms of total energy input, including energy input during grinding. The slurry viscosity is therefore expected to increase along the pipeline and the pumping concentration must be low enough such that the increase in yield stress does not cause transition to laminar flow towards the end of the pipeline. The base viscosity of two ore types is determined using a rotational viscometer for minus 212 micron slurries. Viscosity variations with ore particle sizes of d99 212, 425 and 600microns are investigated. For each of these particle sizes the expected increase in viscosity during pipeline transport is simulated on a laboratory scale by agitation with a turbine mixer in a baffled container to give the same equivalent unit energy input as will occur in the pipeline. Tests were also conducted on the d99425 micron slurry in a 2.7 metre diameter DN 200Wheelstand, rotated for the 23 hour transit time and at a similar relative velocity as in the pipeline. Based on the Wheel stand test results the yield stress is expected to increase by a factor of 3.1during pipeline transit. The preliminary design allows for a reduced pumping concentration to accommodate this increase in yield stress towards the end of the pipeline.

Jan 1, 2007

By J. W. Drelich, H. Zeng, L. Xie, X. Cui, Z. Gao, M. Tian, Y. Hu, W. Hu, W. Sun

Flotation is a mineral separation technology that uses the surface hydrophobic difference of valuable minerals and their associated gangue minerals to separate them. The flotation collector adsorbs on the valuable mineral surface to improve its hydrophobicity, and metal ions are usually used as activators to enhance the collector adsorption. Different mineral surfaces show different charging properties and adsorption behaviours with the collector or metal ions. Moreover, the order of adding metal ions and the collector also show significant and different flotation performance for the same mineral. The adsorption or flotation behaviour or both are governed by mineral surface chemistry, such as surface charging and reactivity. In our recent studies, force measurement using atomic force microscopy (AFM) was employed to study the surface charging and adsorption mechanism of metal ions and collectors. By preparing collector-functionalised AFM tips, AFM can directly measure the adsorption strength of the collector on the mineral surface to determine the anisotropy of mineral surface reactivity. The preparation of collector-functionalised AFM tips provides the foundation for further study of the activation role of metal ion in the collector adsorption. The application of AFM for force measurement in the surface chemistry of mineral flotation can provide important information on developing flotation reagents and lays the foundation for the specific regulation of mineral surface characteristics. Keywords: Scheelite; flotation; surface chemistry; AFM force measurement

Jan 1, 2020

A. H. Mokken: I am pleased to have been given this opportunity to make a contribution to Dr Muller&apos;s paper tonight. The reason for this is that, at one stage in our careers, we were associates in the same undertaking. Dr Muller, then fresh from University, with a degree in pure science, had just stepped onto the first rung of the ladder, which, it was then thought, would lead him to a career in gold extractive metallurgy. However, endowed with an enquiring mind and conscious of a lack of fundamental training in general metallurgy and engineering, he felt a need for further academic study. To meet this need and, more importantly, to meet the necessary finances, he found his opportunity in steel. Armed with a bursary, he bade farewell to gold and proceeded overseas, to the University of Sheffield, to train as a steel metallurgist. The outcome of these academic efforts, which were followed by assignments in the steel industry and a further period at Sheffield, is the man we have listened to tonight-a highly qualified metallurgist who has displayed a sound knowledge of his subject. In choosing Sheffield, Dr Muller became associated with a steelmaking centre of world renown, a centre usually credited with the first systematic production of alloy steels, as far back as the 18th century. Since that time, great advances have been made in the production of alloy steels and this is especially so in the last decade or two, when major developments in civil, mechanical, electrical, aeronautical and nuclear engineering have been made possible by the development of steels with improved properties. In spite of spectacular advances in the technology of non-ferrous alloys, plastics and other materials of construction, steel has maintained its role as a pre-eminent material for engineering use. With the gradual accumulation of data on the properties of steels, and the use of thin film electron microscopy, to study the behaviour and characteristics of such phenomena as dislocations in metals and other microstructural features, the physical metallurgist appears to be approaching the stage of an exact understanding of such phenomena as strength, ductility and brittleness-a knowledge which could lead to close control of such properties and, therefore, to the attainment of the highest goals. An interesting example, illustrating the use of fundamental principles, based on physico-metallurgical research, is the development of the maraging steels developed by Bieber at the International Nickel Company. These steels have met the extreme technological requirements of the space age by providing the material for the cases of large rockets in which qualities, such as high tensile strength, toughness, workability and weldability are most important. Attractive as they might appear to be in considerations of savings of weight, cost of erection, transport of materials and foundations, the use of high strength steels has been accompanied by special problems such as brittle fracture, hydrogen embrittlement, notch-toughness and fatigue. It has been found that high strength steels, which performed acceptably in conventional tension tests, were found to undergo failure, in a brittle manner, in service. Hydrogen embrittlement has caused spectacular failures at a fraction of the normal ultimate tensile strength, and a lack of correlation between fatigue and tensile strengths has diminished the advantages to be obtained from the use of high strength steel in some applications. A new approach to the selection of materials for engineering design has resulted from a consideration of these phenomena in which strength, as such, is no longer as significant as it was previously. Parallel with the development of high strength steel has been the need for suitable techniques for joining component parts and here welding has played an

By D. G. Krige

D. G. Krige: The stimulating contributions from Messrs Cooke, Curtis, Martus, Johnston, Van Wyk, Van Rensburg and Wroth are appreciated by the author and provide clear evidence of the importance not only of the principle of assistance to gold mines, but also of its scope and form. In general all the arguments, raised by Messrs Cooke, Martus and Johnston in favour of extending the scope of the new assistance scheme can be accepted as sound.

By Peter N. Calder, Alan Bauer

The complexity of modem open pit load-haul-dump systems is described, and the need for realistic working models is pointed out. Both standard and probabilistic simulation procedures are discussed, probabilistic simulation being indicated as the more powerful and practical technique. Acquisition of the necessary pit operating data required for simulator construction and general modelling procedures are described. Finally, results from a typical open pit simulator are used to indicate the scope of the technique.

Jan 1, 2014

Copper cyanide is a common component of cyanide-treatable precious metal ores. The copper concentration in production heaps can be predicted from laboratory column tests, but the exact correlation is not necessarily intuitive. Generally, heap leach operators like to keep copper concentrations in solution below 300?500 ppm and may note problems with gold recovery and cyanide consumption when copper concentrations exceed this amount. There are several methods of copper removal from cyanide solutions including ion exchange; direct electrowinning; acidification, volatilization, and recovery (AVR); and sulphide precipitation such as the sulphidization, acidification, recycling, and thickening (SART) process. SART involves acidification with addition of soluble sulphide, separation of the resulting copper sulphide precipitate, and addition of lime to re-establish alkalinity prior to returning the solution to the leaching process, recovering both copper and cyanide as valuable products. In principle SART is very simple. Yet some SART plants that have been built may have been unnecessarily complex. This paper explores the basics of SART and makes the case for a simple plant design as applied to the heap leaching circuit.

Jan 1, 2012

By A. Schippers, S. Hedrich

Chalcopyrite represents the most important copper resource and is currently mainly exploited using pyrometallurgy. Bioleaching is an environmental-friendly and economic alternative to conventional mineral processing technologies. The technical feasibility of stirred tank reactor bioleaching of copper ore concentrates using acidophilic iron-/sulfur-oxidising bacteria is proven, however, often incomplete due to inefficient chalcopyrite dissolution. A decoupling of bacterial growth from bioleaching activity in one bioreactor by adjusting appropriate temperatures for both process steps allows for almost complete chalcopyrite dissolution. Separation of copper and recovery of other metals from the PLS was possible by implementation of different biological and chemical techniques. The Ferredox process has been proposed to treat limonitic laterite ores for cobalt and nickel recovery by means of anaerobic reductive dissolution with autotrophic, acidophilic bacteria. As a modification, aerobic reductive dissolution of laterites with acidophilic bacteria has shown economic and environmental advantages over anaerobic reductive dissolution. The reductive bioleaching technology has the potential to increase metal recovery in existing mines, and transform unexploited ores and limonite stockpiles into valuable resources. Keywords: Biomining, bioleaching, chalcopyrite, laterite, Ferredox process

Jan 1, 2020