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By C. T. Connor, O&apos

The fifth University of Cape Town (UC1)/University of Stelenbosch (US) Minerals Processing Symposium was held at the Faculty of Engineering, University of Stelenbosch on Friday, 8th August, 1986. This series was initiated in 1982, the symposia being held alternately in Cape Town and Stelenbosch. The South African Institute of Mining and Metallurgy has kindly agreed to sponsor this event for the next three years. Cokemaking On the initiative of the European Cokemaking Commitee, the First International Cokemaking Congress will be held in Essen (F.R. Germany) from 13th to 19th September, 1987. It is being organized by the Steinkohlenbergbauverein, together with the Verein Deutscher Eisenhitenleute.

Jan 1, 1986

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 S. Asokan, Ch. V. G. K. Murty, R. Upadhyay

Zircon (ZrSiO4) is the most frequently found zirconium ore occurring in heavy mineral sands, from which it is extracted as a co-product in the production of ilmenite and rutile. Estimated world reserves of zircon are 124 million tonnes, highest in Australia (around 56 million tonnes), followed by Mozambique (15 million tonnes), India (12.6 million tonnes) and South Africa (12million tonnes). Zircon finds its application in ceramics (54%) and refractory industries (14%), which account for 68% of zircon?s total world consumption of 1.20 million tonnes. The rest (32%) is consumed in foundry, TV glass, zirconia chemicals and other applications. It is a long recognized feature of the zircon industry that supply is totally dependent upon the concurrent production of titanium minerals. But in recent years, several new developments such as the discovery of zircon-rich mineral sands in Australia, the continuing trend of higher prices of zircon and increase in trade of lower quality zircon have changed the scenario. These factors have prompted more intensive efforts to increase the recovery of zircon, particularly in the process for separating zircon from alumino-silicate minerals. Tata Steel?s proposed heavy mineral sands deposit contains on an average 10% heavy mineral content and 15% slimes of the size below 63 micron. Mineral assemblages of the resource contain65?70% ilmenite, 4?6% rutile, 4% zircon, 16% sillimanite and other minor minerals. Wide size range and tenacious iron oxide coating are the inherent characteristics of this deposit. High concentration of sillimanite among the non-magnetic heavy minerals and heavy surface coatings on the minerals are challenges for high recoveries and production of premium grade zircon. This paper outlines various process steps that were taken at pilot plant stage to produce zircon of premium grade and subsequently develop the appropriate flow-sheet.

Jan 1, 2007

By M. Cortie, L. Iorio

The amount of nitrogen contained in an austenitic stainless steel affects its properties significantly. The maximum content of nitrogen feasible in a range of experimental low-nickel austenitic stainless steels containing significant additions of copper is evaluated The nitrogen content is calculated based on published interaction coefficients, and by use of the FACT thermodynamic computer program and database. The differences between the two methods of prediction can be attributed to inaccuracies in the published interaction coefficients. The calculated nitrogen contents of some experimental alloys are compared with the actual nitrogen contents obtained when the alloys were produced in an induction furnace. The actual nitrogen contents obtained were higher than those predicted by either method, and this is believed to be due to a super saturation of nitrogen occurring in the melt, which was not relieved during the short period in which the charge was molten. It was evident that the experimental alloys can contain between 0,25 and 0,40 per cent nitrogen (depending on temperature, composition, and nitrogen pressure) when melted in a sealed induction furnace, and this is more than adequate for the purpose of forming and stabilizing the austenite phase.

Jan 1, 1994

By W. Wittke, D. Schröder, H. Polczyk

ABSTRACT: The Urft dam, the Diemel dam and the Eder dam are three old masonry damslocated in Germany. The paper describes the rehabilitation measures and the correspondingstability analyses carried out to adjust the dams to today's standards. Remedial works such as theexcavation of inspection galleries, the installation of a grout curtain with a downstream drainagescreen as well as the installation of pre-stressed permanent tendons were carried out. In all threecases the stability of the upgraded dam could be proven on the basis of finite element analyses. ZUSAMMENFASSUNG: Die Urft-, Diemel- und Edertalsperre sind drei alte deutscheSchwergewichtsmauern aus Bruchsteinmauerwerk. Im Beitrag werden Instandsetzungsmaß-nahmen und Standsicherheitsnachweise beschrieben, die erforderlich waren, um die Staumauernan den heutigen Stand der Technik anzupassen. Hierzu wurden Kontrollgänge aufgefahren,Dichtungsschleier mit luftseitiger Dränage sowie vorgespannte Daueranker hergestellt. In allendrei Fällen konnte unter Berücksichtigung der Instandsetzungsmaßnahmen die Standsicherheit derMauer auf der Grundlage von Berechnungen nach der Methode der finiten Elemente nachgewiesenwerden. RÉSUMÉ: Les barrages de Urft, Diemel et Eder sont trois barrages-poids vieux allemandsconstruit de maçonnerie de moellon. Dans cet article, opérations de remise en état et preuves destabilité sont décrits, qui étaient nécessaires pour adapter les barrages à l'état de la techniqueactuelle. À cet effet, des galeries de contrôle étaient creusés, voiles d'étanchéité avec des drains àla face aval et tirants précontraints permanents étaient fabriqués. Dans tous les trois cas, tenantcompte des opérations de remise en état, la stabilité de la barrage pouvait être prouvée à partir decalculs selon la méthode des éléments finis.

Jan 1, 2003

By J. H. Bulloch

A study of fatigue-crack growth in a rail steel and an associated weld metal showed that fatigue cracks in the deformed rail steel had slightly higher growth rates than in undeformed rail steel. The crack growth in the weld metal was appreciably faster than either and exhibited plateaux that were above the upper bound reported for rail steels, The cyclic-life curves showed that, over an indicated service period of 8 years, both the deformed and the undeformed rail steels should not be prone to fatigue failures under the defined mechanical conditions, For the weld metal, the cyclic-life curves indicated the possibility of premature fatigue failures but, as al the input parameters represented a 'worst case' situation, this assessment could wel be ultra-conservative. In the rail steels at low levels, the extension of the fatigue cracks occurred by ductile striated growth. At a Kmax level of 35 MPa .J/n, cleavage facets formed initially on the surface of the fatigue fractures, and their incidence increased with Kmax until the onset of fast failure, The microstructure of the deformed rail steel exhibited cleavage facets at a slightly lower Kmaxvalue, and contained more cleavage facets at a given K level, than the microstructure of the undeformed rail steel. The average size of the cleavage facets agreed well with the pearlite nodule dimensions of 60 to 100 ILm. The microstructure of the weld metal was much coarser and contained columnar grains. At al f.K levels, the dominant fracture mode was cleavage, with small isolated regions of ductile striated fatigue cracks. The size of the cleavage facets varied from 150 to 600 ILm, and such a variation may be explained by the fact that cracking generally occurs in association with the pro-eutectoid ferite phase, and that this phase is distributed along some boundaries of prior austenite grains and the boundaries of the coarser columnar grains.

Jan 1, 1987

By M. F. Dawson, S. B. Davis

SYNOPSIS This paper describes the size reduction of limestone smaller than 1 mm by attrition grinding, using quartz (1 to 3 mm) as the grinding medium in a laboratory rig. The ore and grinding medium were closely sized fractions, and the particle size of the grinding medium was at least three sizes larger than that of the ore. Batch tests showed that the size reduction of the feed followed two first-order rate processes: one for an apparently fast-breaking ore fraction, and a slower one for most of the ore (86 per cent). The production of fines (material smaller than 10 µm) followed a zero-order rate equation. Continuous runs showed that the process could be modelled as a perfectly mixed system provided that the discharge screen did not selectively retain coarser particles of ore. The ratio of medium to ore (both by mass and by volume) was found to have a significant effect, the most energy¬efficient grinding occurring at ratios between 1 and 2. The solids content of the slurry also had a significant effect, the greatest efficiency being obtained at 50 to 65 per cent solids (by mass). Within these energy-efficient ranges of medium-to-ore and solids content, the lower values resulted in maximum feed breakage and the higher values in maximum fines production. The most important factor in the breakage of feed was the difference in the particle size of the grinding medium and the ore; the amount of ore broken out of the feed size for a given energy input was found to increase linearly with the difference between the size of the grinding medium and that of the ore up to 2 mm. The production of fines increased similarly, but only up to a difference of about 1,2 mm, when it levelled off at about 5 kg/kW ? h. The mechanism proposed for attritioning is mainly the abrasion of 1 µm planes of limestone along the crystal axes, with some fracturing along existing lines of weakness. The energy consumption of nearly 200 kW ? h per ton of product smaller than 10 µm appears to be higher than that in other types of milling. It is concluded that the value of attrition milling lies in its selectivity; in this work, only the softer limestone and virtually none of the hard quartz grinding medium was comminuted.

Jan 1, 1989

By D. Duncanson, R. J. Davidson, C. G. S. Schmidt, G. A. Brown, J. D. Taylor, N. W. Hanf

The development of a cyanidation route for the treatment of gold-plant gravity concentrates is described. This work was carried out as a natural consequence of earlier improvements to the recovery of gold in the gravity sections of existing gold plants. Concern about the environmental aspects of the conventional amalgamation process was further motivation in the development of an alternative procedure. It was found that the partial pressure of oxygen in solution was the most critical reactionrate determining factor. The rate of gold dissolution was shown to be controlled by the reaction rate below approximately 25°C and by diffusion above that temperature. Thus, the efficiency of pulp agitation and oxygen partial pressure permitted a high degree of control over the rate of gold dissolution. The use of oxygen rather than air significantly increased the dissolution rate of both gold and silver. The solubilities of sodium, calcium, and potassium aurocyanide complexes under simulated reaction conditions were studied, and it was established that the values for the sodium and calcium complexes were limited, requiring the use of lower pulp densities and higher reaction temperatures for the treatment of very high-grade concentrates. The presence of excessive amounts of tramp iron in such concentrates of up to 25 per cent by mass was found to result in poor gold dissolution owing to the cementation of gold under certain reaction conditions. The maintenance of high oxygen partial pressures, together with high cyanide concentrations, high pH levels, and the presence of calcium in the system, all aided in passivating the tramp iron. Copper, gold, and iron were shown to be the main cyanide-consuming elements. Optimum gold dissolution was obtained in a closed bench-scale reactor incorporating a flotation-machine type of agitator for effective pulp agitation and oxygenation, the oxygen being recirculated. The reactor was designed to operate under a very small positive pressure. After reaction times of 2 to 6 hours at temperatures in the region of 30°C, both the gold and the silver dissolution averaged 99,S per cent. In the case of very high-grade concentrates, it was shown that gold could be floated in the same reactor by drawing off the flotation concentrate through a port in the reactor. A flotation recovery of 87,4 per cent of the gold in 3,9 per cent by mass was obtained. The recovery of osmiridium from cyanide-leach residues and the electrowinning of gold from the pregnant liquors were also considered as processing aspects in the treatment of gold-plant belt concentrates. Final table tailings with a gold concentration of 10to 12 gft were obtained. The results of the investigation strongly suggested that an alternative cyanidation route for the treatment of gold-plant concentrates was technically feasible, provided the system included effective agitation and oxygenation. In the case of lower-grade concentrates, cyanidation followed by the electrowinning of gold appears to be the simplest route, while, for high-grade concentrates, flotation and smelting of the concentrate followed by cyanidation of the flotation tailings are suggested. The use of oxygen rather than air for sparging the pulp would depend largely on the grade of material to be treated. A preliminary comparison of the costs related to the amalgamation and cyanidation processes for the recovery of gold from such concentrates indicated very similar capital and operating costs. It was concluded that other considerations, e.g. toxicity, safety, and lower labour requirements, were more important than such small cost differences. Furthermore, the 'opening' of the gold-plant gravity circuit, which would be made possible by a cyanidation route, would further reduce the operating costs because of the subsequent decrease in the tonnage to be treated.

Jan 1, 1978

By C. G. Verkerk

The paper discusses the transport and rheological characteristics of high concentration slurries or pastes for backfill purposes. The influence of pipe size, particle size distribution, additives, coarse material and mass concentration on these characteristics is addressed.

Jan 1, 1988

By C. G. Schmidt, N. W. Hanf

Many of the pyrite concentrates produced at South African gold and uranium mines contain economically significant amounts of nickel, copper, and cobalt, apart from gold and uranium. In some cases these concentrates are leached for uranium, followed by roasting and cyanidation of the calcine for the extraction of the gold. Even if their intrinsic value is excluded, the presence of appreciable amounts of base metals can lead to excessive consumption of cyanide and zinc in the conventional process for the recovery of gold from calcines. The extent to which base metals, particularly cyanocides in the calcine, can be made soluble in water or in dilute acid by variation of the pyrite roasting conditions was investigated. An important finding of the investigation was that the conditions producing favourable base-metal sulphation (and hence solubility in water) also produced good sulphation of the uranium values that had not been extracted in a conventional leach of the original pyrite. Thus, if the pyrite is subjected to a sulphating roast followed by quenching in water (or dilute acid), it should be possible for the uranium to be extracted in improved yield. The need for leaching of the pyrite would therefore fall away. It was noted that sulphation roasting also resulted in improved gold extractions when the calcine was subsequently subjected to conventional cyanide-extraction procedures. From well-sulphated calcine, the dissolution of base metals and uranium appeared to be fairly independent of leaching pH, time, and temperature. On an inadequately sulphated calcine, Iow pH values during leaching gave increased extraction of copper and uranium but did not have a significant effect on the dissolution of nickel and cobalt. The extraction of uranium increased with increasing leaching time and temperature, while only longer times at a Iow pH value markedly improved the extraction of copper. The extraction of cobalt was fairly independent of leaching time and temperature, but the results were insufficient for a conclusive evaluation to be made of their effects on the dissolution of nickel. A cursory economic assessment 'indicated that sulphation roasting applied to the pyrites investigated is likely to be very profitable.

Jan 1, 1979

By R. Tabanrad

This article examines the stability analysis of intake tunnels of Karun III dam & power plant. The tunnels have been excavated with a 3-dimensional and fairly complicated geometry in Pabdeh weak formation rocks at Zagros region. Due to the particular excavation geometry and inapplicability of plane-strain status, the stability of these excavations has been analyzed using both the 3D elastic software Examine 3D and the 2D non-elastic software UDEC simultaneously. The geomechanical parameters have been estimated by back analyzing the measured displacements from instruments, which were installed in the tunnels, and the parameters have been used in non-elastic numerical model. The results of the analysis show that the strains occurred in surrounding rock mass are less than the allowable failure strain and that the tunnels are stable. In addition, considering the results of non-elastic 2D analysis, the areas with unstable potentiality were identified and after the instability occurred at the areas predicted by the model , it was stopped up to the allowable strain limit through installing additional support . Bei diesem Artikel handelt es sich um eine Analyse der Festigkeit der Eingangstunnel des Damms und Elektrizitätswerks “Karoon III”.Die Tunnel haben drei Dimensionen und eine sehr schwierige Form und sind in brüchigen Felsen in “Pabedeh ,in Zagrossgebirge”eingebohrt.Aufgrund besonderer Form der Bohrung wurde ihre Festigkeit durch elastische dreidimensionale Soft-ware analysiert,die gleichzeitig die drei-und zweidimensionale unelastische Soft-ware UDEC benutzt hat.Die geometrische Parameter sind mit Ausnützung der Analyse berechneter eingesetzter Versorgungen,die in Tunnel aufgebaut sind,bewertet.Und diese Parameter sind in Zahlen-Analyse benutzt.Die Ergebnisse dieser Analyse zeigen,dass die Drücke an Felsen weniger als problematischer Grad der Drücke sind und die Tunnel fest sind.Ausserdem wurde mit einem Einblick auf die Ergebnisse der unelastischen zweidimensionale Analyse die eventuellen unfesten Teile bestimmt.Nach entstandener Unfestigkeit an diesen Teilen wurde die problematischeDrücke für Aufbau mehrere Anlagen gestoppt. Cet article examine l'analyse de la stabilité de tunnels de la prise de Karoun barrage III et centrale électrique. Les tunnels ont été excavés avec une 3 dimensionnel et compliqué géométrie dans Pabedeh que la formation faible balance à région Zagros équitablement. Dû à la géométrie de l'excavation particulière et inapplicabilite de statut de tension plane, la stabilité de ces excavations a été analysée en utilisant les deux le 3D logiciel élastique Examen 3D et le 2D logiciel non - élastique UDEC simultanément. Les paramètres du geomechanical ont été estimés en analysant le déplacement mesuré d'instruments qui ont été installés dans les tunnels et les paramètres en arrière a été utilisé dans les modèles numériques non - élastique. Les résultats de l'analyse montrent que les tensions se sont produites dans masse du roc environnante est plus petit que la tension de l'échec admissible et que les tunnels sont stables. De plus, les régions avec potentialité instable ont été identifiées étant donné les résultats de 2D analyse non - élastique, et après que l'instabilité se soit produite aux régions prédites par le modèle, il a été arrêté jusqu'à la limite de la tension admissible à travers installer du support supplémentaire.

Jan 1, 2003

By Q. P. Campbell

Activated carbons used in the precious metals extraction industry are characterized by large internal surface areas and a great affinity for metal ions. The purpose of this research is to prepare activated carbon from a South African bituminous coal by physical activation that is suitable and cost-effective for use in the extraction of metals. The quality of the coal-based activated carbon may not prove to be as good as activated carbon produced from other traditional sources, but the production costs involved may make South African coal a feasible alternative feedstock. The activated carbons produced were characterized by Brunauer, Emmett, and Teller (BET) surface area, activated carbon pH, and phenol adsorption studies and the results compared to the results from a commercially available activated carbon, Norit RO 0.8 (control sample). Bituminous coals from various sources including Witbank Seam 4 and Free State coal were used in this study. The preparation method chosen was physical activation using superheated steam. The effects of process variables such as activation time (1-3 h) and temperature (600-800°C) were studied in order to optimize those parameters. The activated carbon surface area was characterized by means of nitrogen adsorption isotherms at 77K. BET surface area analysis showed that Witbank Seam 4 coal activated at a temperature of 800°C and activation time of 3 hours resulted in a surface area of 340 m2/g. Quality control of each sample was performed by measuring the pH of a known amount of the prepared activated carbon in distilled water over time. Results showed that the pH of some of the prepared activated carbons reached a value of 11. Phenol adsorption results for the different activated carbons prepared corresponded well to the results obtained for the Norit RO 0.8 activated carbon sample.

Jan 1, 2012

By C. R. S. Needes, M. J. Nicol, G. T. W. Ormrod, N. P. Finkelstein

The industrial application of a dioxide electrode uranium by C. R. S. NEEDES, Ph.D. (London) (Visitor), M. J. NICOL, Ph.D. (Rand) (Visitor), N. P. FINKELSTEIN M.Sc. (Rhodes), Ph.D. (Liverpool) (Fellow), and G. T. W. ORMROD (Visitor) A correlation between the potential of a U02 electrode and the rate of recovery of uranium has been proved in laboratory and plant trials. When the recovery rates change because of variation in the concentrations of Fe(III), Fe(II), SO4 'and H+, a positive correlation is observed. However, an 2- increase in the concentration of phosphate in solution produces an increase in the U02 electrode potential but a decrease in the rate of leaching of U02. The correlation between the U02 electrode potential and the rate of leaching of U02 is then negative. It is concluded that, as a control device, the electrode cannot compete with the platinum electrode for use on certain plants. Nevertheless, the U02 electrode will act as a useful warning device if the total concentration of iron in solution decreases to below a level concomitant with the economic recovery of uranium. Furthermore, because of the positive correlation between the U02 electrode potential and the phosphte concentration, the electrode will also be of value in the detection of an increase in the phosphate level in solution. When it was incorporated in a suitable industrial probe, the electrode was found to be able to withstand the rigours of the leaching conditions in a large pilot-plant pachuca, and only failed after six weeks' operation.

Jan 11, 1975

By A. M. Copeland, A. J. Barrett

The use of hydraulically placed backfill in the deep gold mines of South Africa has increased rapidly over the past decade. The backfill that is delivered to the placement areas has a relatively high moisture content, and it is therefore pumped into geotextile containment bags that allow excess water to drain out but retain solid particles. The specific problems associated with this process that are addressed in this paper include the loss of solid particles entrained in the excess water, the rate of strength gain of the backfill, and the post-placement shrinkage of the backfill. The results of laboratory, surface stope, and underground tests are described. The parameters that were varied included the slurry density, the particle-size distribution, and the feed rate of the slurry the concentration of added flocculant; and the type of geotextile used as containment. Suggestions are given for the optimization of the placement properties of backfill.

Jan 1, 1994

The support of stopes by means of yielding timber props, with or without hydraulic props, is common in the mining industry. This paper discusses the ability of these support systems to absorb energy during a rockburst, and further investigates their capability to withstand a range of ground velocities. It is found that the rate of stope closure prior to a rockburst strongly influences the behaviour of these support systems during the rockburst. It is recommended that hydraulic props be used in conjunction with yielding timber props in these conditions, and that yielding timber props with an increased yield range are desirable.

Jan 1, 1991

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

By A. G. W. Knock

Logistic curves are used in the prediction of the primary production of tin metal from estimated worldwide resources. Such curves are asymptotic to the total cumulative reserves, which are a function of prevailing metal prices and total resources. The figures for tin resources are based on the findings of Robertson and of Brobst and Pratt, the latter authors including estimates of undiscovered deposits. It is suggested that tin production will slump from the present level to 100000 long tons per annum by the end of the century, but will reach a new peak of 200 000 long tons some thirty years later. It appears that production will never exceed 300000 long tons per annum unless substantial new resources, excluding those predicted, are discovered.

Jan 1, 1979

By Nareetsile S., J. Kabwe

In the past few years Orapa and Letlhakane Mines have investigated various new technologies to enhance and improve production. The goal has been to improve productivity, reduce operating costs and make the mines more profitable The advent of Global Positioning Systems (GPS) in the early ’80s brought about the realization of this concept. Orapa and Letlhakane Mines started using GPS in 1999 and it is now used atLetlhakane mine for surveying, vehicle tracking and dispatching and for floor control system. Orapa Mine uses GPS in surveying, floor control and drill navigation. It will also be used for vehicle tracking and dispatching in mid-2001 following the installation of a fleet management and control system. The introduction of GPS at both Orapa and Letlhakane mines have improved productivity and reduced operating costs in the areas of drilling, topographic survey and floor control. In vehicle tracking and dispatching GPS has allowed for real-time information management for timeous decision-making which optimized fleet run time hours.

Sep 1, 2001

By C. Rodrigues

This investigation compares the gas adsorption behaviour and capacity of three bituminous coals from South Africa, of similar rank, by assessing the characteristics of the raw coal, as well as the resulting float and sink fractions (at 1.80 cm3/g) obtained by density separation of crushed coal samples. Calculations were also made to obtain the raw coal gas storage capacity from the weighted contribution of both float and sink fractions results, thereby permitting comparison with the analysed results of the raw coal. The study demonstrated that the clean fraction of a coal has the highest capacity to retain gas in the sorbed state, followed by raw coal, and lastly the sink fraction, and re?confirmed previous investigations that showed minerals to be inhibitors of gas adsorption and retention.

Jan 1, 2008

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