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By H. M. Wells

The influence of economics on the design of mine shaft systems by H, M. WEI.LS, Pr. Eng., M.Se. (Fellow) The purpose of this paper is to review some of the more important economic criteria that must be taken into account when planning the number, type, size, and distribution of the shafts, or shaft systems, required to exploit a block of payable ore. The postulated philosophy governing the selection of design parameters is that of optimization of the financial return per rand of venture capital employed for the project as a whole. While this criterion may not necessarily satisfy the investment objective under all circumstances (as, for instance, where risk limitation is important), it does serve to place the relative economic advantages and disadvantages of specific design features, which may sometimes be in opposition, into correct perspective with respect to the overall economics of the project under the prevailing economic climate at the time of the investigation.

By K. G. Tshilombo

It should be possible to use an oxygen probe to follow iron removal during converting, if local equilibrium is established between iron in the matte and the iron oxides in the fayalitic slag. The oxygen activity is expected to increase strongly as the iron content of the matte decreases (and the iron activity in the matte drops) while the iron oxide activity in the slag remains constant. To test this idea, oxygen activities were measured in synthetic mattes (made up of mixture of Cu2S, FeS and Ni3S2) when in contact with silica saturated SiO2-FeO slags, and silica and (initially) magnetite saturated SiO2-FeO-Fe3O4slags. The measurements were performed on laboratory scale at 1250°C, using an oxygen probe consisting of a stabilized zirconia electrolyte and Fe/FeO reference, and containing the sample in a silica crucible. Electrical contact to the matte was by means of a platinum wire. The measured oxygen activity was sensitive to the iron content of the matte, and insensitive to the composition of the gas atmosphere over the melt. The measured oxygen activity was generally close to that predicted by Fact Sage calculations. This indicates that such oxygen activity measurements could be useful to monitor iron removal during converting of nickel-copper-iron mattes.

Jan 1, 2006

By A. A. Bradley, K. H. Stanton, P. J. D. Lloyd

A theory for the dynamic balancing of centrifugal mills is presented that takes account of the mass and shape of the charge in a gyrated grinding tube and the way in which the charge lags behind the centrifugal force vector. predicts that a corresponding lag in the angular position of the counterweight centres of mass is necessary, and that the fractional filling should be increased if the ratio of gyration diameter to tube diameter is decreased. A worked example is given for a I MW centrifugal mill that was balanced adequately by the methods described.

Jan 1, 1983

By Z. T. Bieniawski

Estimating the strength of rock materials by Z. T. BIENIAWSKI, Pr.Eng., D.Sc.(Eng.) (Visitor) A practical approach for estimating the strength behaviour of rock materials is presented. The demand for data on rock properties in engineering design is considered, and it is shown that, based on the trends observed in the South African mining industry during the last twenty years, simple and easy-touse methods for the estimation of the uniaxial and triaxial strength of rock materials are needed. It is shown that the uniaxial compressive strength of rock can be conveniently determined from the pointload strength index, which is obtained underground on unprepared rock cores. The triaxial strength of rock can be estimated for most practical purposes from empirical strength criteria. Two such criteria are proposed that allow estimation of the triaxial strength for rock materials to about 10 per cent. The only input required for these criteria is the uniaxial compressive strength. The validity of the practical approach outlined in this paper is confirmed from the experimental tests conducted on some 700 rock specimens representing five rock types. A practical example for the prediction of rock strength is given.

Jan 3, 1974

The AK6 kimberlite is situated 25 km south of the Debswana Orapa Mine in Botswana and was discovered by De Beers geologists in 1969 during the follow-up of geophysical targets in the Orapa area. The kimberlite was not extensively pursued at the time as the initial bulk sampling indicated it to be of limited size and low grade, factors largely contributed to by the basalt breccia capping. Completion of high resolution integrated geophysical techniques and drill bulk sampling to depth recovered 97 tons of kimberlite during 2003 and 2004, which led to the increased size and grade estimates. Bulk sampling by Large Diameter Drilling (LDD, 23 inch diameter) commenced in 2005; 13 holes were drilled to a cumulative depth of 3,699 m and 689 carats of diamonds were recovered. In July 2006 the De Beers Mineral Resource Classification Committee classified these Phase I LOO results at a High Inferred level with an average grade of 24 carats per hundred tonnes (cpht) at a bottom cut-off of +1 mm, and a modeled average diamond value of 150 dollars per carat. A second phase of LDO drilling was initiated in 2006, and bulk sampling by trenching commenced in 2007 in order to deliver a resource estimate at indicated level. An Indicated Resource of 11.1 million carats at an average grade of 22 cpht was declared for the deposit mining lease application lodged in 2007.

Jan 1, 2010

By T. Könighofer

Processing of copper-cobalt orebodies in Africa?s Copper Belt has received much recent attention with various flowsheet options being developed specifically for the refining of cobalt. The upfront leach and copper refining routes are well understood whereas the refining of cobalt is more complex, requiring numerous impurity removal steps to produce high purity metal. This paper describes an investigation into one of the possible processing steps in the refining of cobalt, namely, cobalt solvent extraction. The paper emphasizes the importance of upstream impurity removal to achieve the required cobalt solvent extraction feed composition. 18% (v/v) Cyanex 272 in an aliphatic diluents is used as the organic phase. The pH profiles in the various stages are evaluated in order to obtain a raffinate containing <10mg/l cobalt and with a low magnesium content reporting to the stripping section. The pH profile is also used to minimize the impurity deportment to the cobalt electrolyte. The effect of adding a pre-neutralization stage before feeding the organic phase to the extraction circuit is investigated. Zinc build-up in the stripping stages is also looked at. It is recommended for the purpose of this study that four extraction stages at an O:A ratio of 1 with a pH profile from 4.9 to 5.7 be employed. The pH is controlled with 40g/l NaOH. A pre-neutralization stage is required, where the organic phase is contacted with 10 M NaOH. Two scrubbing stages are recommended at an O:A ratio of 40, using the cobalt electrowinning advance electrolyte as the scrub liquor. Three strip stages are to be employed at an O:A ratio of 0.67.

Jan 1, 2009

By L. A. Cornish, B. Eatwell

Austempered ductile iron (ADI) is a ductile iron which has had its matrix toughened by an austempering process. The alloy exhibits remarkable properties, which it owes to a unique microstructure comprising graphite nodules embedded in a matrix of acicular ferrite and carbon-enriched austenite. The properties include high strength combined with high toughness and ductility. Excellent wear resistance is also exhibited because, under strain, the austenite is transformed locally to martensite. Samples of ductile iron were austempered to produce grades I and IV AD!, and were tested at various strain rates. The grade IV ADI exhibited higher strengths and lower difficulties, and the strengths were found to be more dependent on strain rate. Transformation to martensite was more prevalent at the lower strain rates. A mechanism of increased flow stress and work hardening in the austenite of the matrix is proposed to explain the higher strength of the samples at high strain rates, where significantly less martensite is formed in the matrix.

Jan 1, 1995

By J. A. Cruise

From 27th to 29th September, The South African Institute of Mining and Metallurgy held a Colloquium on Equipment Alternatives in Underground Mining in Johannesburg. The Colloquium, which was highly successful, was held concurrently with the 1988 Electra Mining Exhibition. Particle characterization The Fourth European Symposium on Particle Characterization is to be held in Nirnberg, West Germany, from 19th to 21st April, 1989.

Jan 1, 1988

By F. Shaw, E. C. Chemaly, A. F. S. Schoukens

Steel-plant dusts are hazardous waste materials, containing toxic constituents such as lead, cadmium, and hexavalent chromium. These dusts can be treated by a number of thermal processes, which are described as used in Western Europe and the USA. A detailed description is then given of the thermal Enviroplas process, which as developed at Mintek and produces disposable slag and valuable byproducts such as ferroalloys containing chromium and nickel, and fumes high in zinc oxide.

Jan 1, 1993

By More-O&apos, R. C. Ferrall, J. W. Wilson

Discussion: M. D. G. Salamon* (Fellow): The theoretical principles behind the techniques of calculation used by the authors was formulated and published during the period 1962- 65. These principles, together with their practical implications, were summarized in a review paper in 1966.1 Now, in February, 1970, we have this long and comprehensive paper discussing the practical application of the theory. I think this is good progress. The large number of practical examples treated in the paper clearly demonstrates not only the skill and knowledge of the authors, but also the versatility of the techniques of analysis. I feel that the techniques described by the authors, and their improved versions, will revolutionize mine planning. I suggest that no shaft system or stoping layout should be planned in future without an adequate rock mechanics analysis. It seems that it will be progressively more difficult to white wash mistakes in rock mechanics and attribute their costly outcome to normal mining risk. We should not, however, become complacent. There is still a lot to learn and do in the future. I do not want to discuss tonight the question of further research but instead, I will raise two other problems which may hinder the future progress of practical rock mechanics. To carry out work similar to that described in the paper, mine managements require men who are capable of doing the work and who have the most efficient techniques of analysis at their disposal. As a research engineer I have been associated with the development of techniques of analysis. I can assure you that these have undergone an amazingly rapid development during the last six years. We have seen the change from the use of the electrolytic tank analogue and manual integration to the method described by the authors, that is, to the employment of the automatic network analogue and computer integration. But this is not the end. In a recent paper Prof R. P. Plewman2 and his co-authors described a completely digital technique. Here the process of calculating on- and off-reef quantities is integrated into one computer run. But these sophisticated methods of computation are useless without men to make intelligent use of them. During the last few years a new breed of mining engineers has appeared on the scene - the Rock Mechanics Engineer. I feel it is timely to call the attention of the senior members of this Institute to the fact that this new breed will succeed in carrying out its duties in a valuable manner only if we manage to attract to its ranks young engineers of the right quality. The requirements are high. The top rock mechanics engineers have to be good mine planners and they have to understand the essence of sophisticated theories. Without these attributes they will not succeed in the long term and the industry will not gain the benefit of available knowledge. To attract this quality of engineers the industry will have to ensure that rock mechanics is accepted by all concerned as a career which could lead to the top echelons of our profession. Finally, I would like to make one or two remarks concerning Part I of the paper. In Sections 3.1 and 3.2 the authors describe methods of determining the normal stress on the reef plane and the convergence distribution in the excavations. The description as given applies only to a horizontal reef. When the reef is inclined the situation is more complex since a shear component of stress on the intact reef and a ride component in the excavations must then also be determined. These can be calculated by methods similar to that described in the paper.3 The practical application of the method of calculating the released energy (Section 3.4) is in the comparison of various stoping layouts to establish an order of preference in terms of decreasing rates of energy release. To carry out these comparisons effectively, the calculations must be carried out in a manner by which the values of energy per unit area obtained are comparable. I would like to note in this respect that the method of calculation given in the paper is valid without reservation only if there is no elastic contact between the hanging and foot walls in the mined-out area. If there is contact then the calculated energy values are correct only if the mined out area is increased in small steps during the analysis. My second point in this regard is concerned with the practical method of calculation as described in Sections 3.4 (ii)-(iii). Firstly, I would like to suggest that the energy calculation should be carried out, as far as possible, by using always the same scale on the analogue. Secondly, the reading of current on a pin should be followed by the removal of that pin to obtain the corresponding reading of potential. The product of these two values will be proportional to the energy released during the mining of the small area corresponding to that pin. The method suggested by the authors tends to mask possible danger points in the layouts, since they obtain an average value for a time period, say, for six months.

By L. Mabiza

Keywords: Pyrometallurgy, circular furnace, PGM, platinum, Peirce-Smith converter, Zimbabwe The Zimplats Smelter, located on the northern part of the Great Dyke, at the Selous Metallurgical Complex, is the only PGM smelter in Zimbabwe. Zimplats? operations currently consist of the Ngezi Platinum Mine, which is a 1.5 Mt/a open-cast and 0.7 Mt/a underground mine, and a Concentrator and Smelter to which ore is transported 80 km by road. The final converter matte product is transported to Impala Refining Services in South Africa for refining and marketing. The Zimplats Smelter operations consist of a concentrate filtration plant, a Drytech flash dryer, a modified Elkem / Hatch 13.5 MVA circular furnace, two 10? x 15? Peirce-Smith converters, and a matte granulation plant. The Zimplats smelter uses a three-field electrostatic precipitator, supplied by ELB-Brandt, for removal of particulates in the gas streams from both the converter and furnace. Construction of the smelter commenced in 1995, under the BHP / Delta Gold of Australia joint venture. The furnace was lined mid-1996, and smelting operations commenced in 1997. Due to design challenges, the Elkem furnace was modified by Hatch in 1998, and the smelter had not reached full capacity when operations were suspended in 1999. Re-commissioning took place in January 2002.

Jan 1, 2006

By FT Cawood

The latest Draft MPRRB was released for public comment in October 2006. The main purpose of the Bill is to collect royalties at mines developed in terms of new-order mineral rights granted through the MPRDA process. The transition phase stipulated in the MPRDA will be completed by 1 May 2009, on which date mineral royalties will become payable to the state. The royalty rates stated in the MPRRB differ from mineral to mineral and also depend on the degree of refinement. Gold is the subject of this paper, where the royalty rate ranges from 1.5% (for refined gold) to 3% (for unrefined gold) of sales revenue. This royalty is payable regardless of profit, with a maximum 75% relief for loss-making mines. The purpose of this paper is to establish whether or not it would be necessary for gold mines to adjust their mineral reserve statements as a result of this Bill. The methodology is to add the royalty costs stipulated in the MPRRB to the actual quarterly information for the past ten years as reported by the gold-mine members of the COM. The rationale is to investigate the impact this Bill (in its current form) would have had on gold reserves that must be left unmined in the ground because of a higher pay limit. The results of the study point to a significant impact on reserves available for mining with the subsequent shortening of mine lives and retrenchment of workers. Had the MPRRB been effective for the 10 years prior to 1996, it would have resulted in 46 tonnes of gold being sterilised. Stated alternatively, 30 700 workers at COM gold mines alone would have lost their livelihoods.

Jan 1, 2007

By More-O&apos, R. C. Ferrall, J. W. Wilson

SYNOPSIS The theory of elasticity has been shown to provide a means of determining the stresses and displacements induced by mining in deep level hard rock mines. This theory has been incorporated into an electrical resistance analogue on which the plane of a reef or ore-body can be modelled to enable the rapid determination of stresses and displacements at points in the plane of the reef. This electrical analogue has been used to determine energy release rates caused by scattered mining, which, to some extent, is a measure of the degree of hangingwall fracture. A correlation between predicted energy release rates and actual underground conditions in stopes in the O.F.S. goldfields has been encouraging enough to define parameters on which future stoping layouts can be based. The use of the elastic theory through the analogue has been extended to the assessment of the influence of stoping operations on off-reef excavations. Underground conditions at damaged and undamaged areas have been compared with predicted elastic stresses at the positions under review, resulting in the derivation of design parameters for use in siting off-reef excavations in the prevailing geological conditions in the O.F.S. and Klerksdorp goldfields. The analogue has also been used to assess the vulnerability of certain un mined dykes to bursting. Correlations between stresses induced on dykes by stoping and dyke behaviour underground has enabled critical stress levels to be determined for several persistent dykes occurring in the O.F.S. and Klerksdorp goldfields. A number of long and short term planning projects at mines operating in the Anglo American Corporation and General Mining and Finance Groups are described, where the data derived from the electrical resistance analogue is processed by computer programmes and assessed on the strength of the design factors derived from the correlation of field work and the elastic theory.

Jan 1, 1970

By Klaus Hasselwander

Fluctuating SO2 gas streams represent a difficult task for designers of SO2 removal systems (with respect to designing the plant), but in spite of the designers great efforts to solve the problems, the operation of such plants requires much care and attention. A fluctuating gas flow usually results from metallurgical smelting processes. Both gas composition and gas flow rate may vary. The first part of this paper outlines some existing technical solutions and presents the Pros and Cons of the different solutions. These solutions outline are: - Sulphuric acid plants operated with a preheater permanent in operation - Sulphuric acid plants with concentration increase of SO2 gas either using regenerative SO2 removal or cryogenic partial condensation of the SO2 - Sulphuric acid plants with additional burning of sulfur to increase SO2 content - Single-stage sulphuric acid plants followed by a tail gas system (which is) designed for periods of low SO2 input In the second part, this paper describes some existing technical solutions for removal of SO2 at low concentrations (i. e. less than 0.1 - 2 %-by volume), which could be used as tail gas treatment system for acid plants. While selecting an appropriate solution, special attention should be given to operating costs, which usually to a large extend depends on the type of absorption agent and the costs for disposal of the sulfur compound produced.

Jan 1, 2009

By Klaus Hasselwander

Fluctuating SO2 gas streams represent a difficult task for designers of SO2 removal systems (with respect to designing the plant), but in spite of the designers great efforts to solve the problems, the operation of such plants requires much care and attention. A fluctuating gas flow usually results from metallurgical smelting processes. Both gas composition and gas flow rate may vary. The first part of this paper outlines some existing technical solutions and presents the Pros and Cons of the different solutions. These solutions outline are: - Sulphuric acid plants operated with a preheater permanent in operation - Sulphuric acid plants with concentration increase of SO2 gas either using regenerative SO2 removal or cryogenic partial condensation of the SO2 - Sulphuric acid plants with additional burning of sulfur to increase SO2 content - Single-stage sulphuric acid plants followed by a tail gas system (which is) designed for periods of low SO2 input In the second part, this paper describes some existing technical solutions for removal of SO2 at low concentrations (i. e. less than 0.1 - 2 %-by volume), which could be used as tail gas treatment system for acid plants. While selecting an appropriate solution, special attention should be given to operating costs, which usually to a large extend depends on the type of absorption agent and the costs for disposal of the sulfur compound produced.

Jan 1, 2009

By M. F. Handley

This paper covers in situ stresses in the Earth?s crust prior to any man-made disturbances, such as mining. It introduces the Southern African Stress Database, which contains primitive stress measurements obtained from locations spread all over southern Africa. The measured stress data shows large variability without any visible trends, except a relationship with increasing depth. The stress database is reviewed briefly, establishing means to measure the variability or dispersion in the measurements, and showing that the dispersion is not as much a result of experimental error as it is a feature of primitive stress. The paper demonstrates from the beginning that the state of stress in rock is highly variable, but that there are well-defined maximum and minimum limits to all the stress components in rock. Formal error analysis is introduced to check the consistency of the database and to separate out a database of consistent stress measurements for use in a primitive stress model. The aim is to provide a picture of primitive crustal stress based on objective stress measurements together with interpretations of how the primitive stress can be affected by the five main influences; namely, depth, rock mass properties, tectonism, isostacy, and erosion. Four elementary models for primitive stress are introduced and compared with the measured data. It is quite clear that none of the elementary models is sufficient to describe the data. In the absence of a better model, this paper suggests a generic model based on the Hoek-Brown failure criterion and the consistent stress database, since it incorporates the variability in the stress tensor that is likely to be encountered underground all over southern Africa. Rock engineers should take every opportunity to obtain local primitive stress data at every mining operation and civil engineering project, and to adjust the proposed model accordingly Keywords primitive stress, Southern African Stress Database, error analysis, generic primitive stress model.

Jun 1, 2013

By C. Musingwini

The South African platinum mining industry is using a number of different mining methods, including variations of the same method, to extract the narrow reef tabular platinum deposits of the Bushveld Complex. These mining methods fall into three broad categories, namely conventional, mechanised and hybrid mining. A question sometimes asked in the industry is whether mechanized mining methods are more efficient than conventional mining methods. An objective answer requires the methods to be evaluated against multiple criteria simultaneously, whereby each criterion has a relative degree of importance in the overall decision. The most efficient method is the one that scores highest on each criterion. However, some of the criteria can be conflicting, such as by increasing dilution the shaft head grade decreases. The analytic hierarchy process (AHP) methodology was selected for the study because it is used to solve problems of this nature. A survey was carried out to determine the relative importance of each efficiency criterion by drawing on the knowledge and experience of mine technical services and project management practitioners in the industry. Efficiency data for four different mining methods drawn from the Egerton (2004) study were used as a case study. The conventional mining method ranked as the most efficient mining method from the four methods considered. The exercise indicates potential for the AHP to be used in the South African platinum mining industry as a tool for selecting optimal layout designs, conduct regular evaluation of the performance of production shafts, or objectively evaluate line managers for promotion. The work described in this paper is part of the methodology used in a current PhD research study at the University of the Witwatersrand. Keywords: Analytic hierarchy process (AHP); multiple criteria decision analysis (MCDA); decision-making; conventional mining, hybrid mining, mechanized mining.

Jan 1, 2008

By I. Gil, J. -C. Roegiers

Abstract Hydraulic fracturing is a technology that has been utilized for more than 50 years in the oil industry. It was originally used for stimulating hard, brittle formations which typically exhibit low permeabilities and roughly linear elastic behavior. Nonetheless, an increasingly important segment of the industry is currently stimulating very soft and poorly consolidated formations; where the assumptions of ideal elasticity and relatively small fluid leakoff fail to hold (e.g. Gulf of Mexico, West Africa, Alaska, East China). In these rocks, hydraulic fracturing stimulation has been used for several years to control and solve critical production problems such as sanding and formation damage (caused during completion and drilling operations). Although successful1, the application of this method has not been accompanied by robust fracturing modeling techniques. In addition, long term production increases were often not materialized. In other words, engineers have been able, via trial-and-error, to avoid major ‘operational’ problems but with no optimization of such treatments. Thus, most hydraulic fracturing projects carried out in poorly consolidated rocks render unexpected results. A recent worldwide survey on net pressures by the Delft Fracturing Consortium (Papanastasiou, 1997) indicated that net pressures encountered in the field are 50% to 100% higher than the net pressures predicted by conventional fracturing simulators. The difference is yet higher for the case of poorly consolidated formations (Pak, 1997). The theory for the construction of a fully coupled elastoplastic three-dimensional hydrofrac simulator, specifically conceived for unconsolidated rocks, is presented in this paper. This novel model calculates the dimensions and changes in shape of the fracture(s) being created during hydraulic fracturing operations. No assumptions are made regarding either shape or geometry of the crack(s). Three main failure criteria are used simultaneously in this model: tensile and shear stresses, and volumetric strain. Incremental forms of the governing equations of equilibrium and fluid flow are presented. The scheme chosen for solving this problem is the finite elements method, which allows both equations to be solved simultaneously at any instant in time. Résumé La fracturation hydraulique est une technologie qu’a été utilisé pour plus que 50 ans dans l&apos;industrie pétrolière. Elle a été utilisée au début pour stimuler les formations dure et fragile que typiquement ont des perméabilités basses et un comportement élastique à peu près linéaire. Néanmoins, un segment plus important de l&apos;industrie stimule actuellement des formations mal consolidées; où les suppositions d&apos;élasticité idéale et de relativement petite fuite de fluide ne se tiennent pas raison (par exemple le Golfe de Mexique, l&apos;Afrique d&apos;ouest, Alaska, la Chine de l&apos;est). Dans ces roches, la stimulation avec fracturation hydraulique a été utilisée pour plusieurs années pour contrôler et résoudre les problèmes critiques de production comme les dommages de formation et d’ensablement (causé pendant les opérations de l&apos;accomplissement et forage). Bien que prospère, l&apos;application de cette méthode n&apos;a pas été accompagnée de techniques robustes pour le modelage de fractures. En outre, les augmentations à long terme de production n&apos;ont pas été souvent concrétisées. En d&apos;autre termes, les ingénieurs ont pu, via le procès et l’erreur, éviter les majeur problèmes ‘opérationnel’ mais avec aucune optimisation de tels traitements. Ainsi, la plupart des projets de fracturation hydraulique ont été exécuté dans des roches mal consolidées et ont rendu résultat totalement inattendu. Une récente étude mondiale sur les pressions nettes par le Consortium Delft pour Fracturer (Papanastasiou, 1997) a indiqué que les pressions nettes qu’ont été rencontré dans le champ sont environ 50% à 100% plus hautes que les pressions nettes prédites par les simulateurs fracturant conventionnels. La différence est pourtant plus importante pour le cas de formations mal consolidées (Pak, 1997). La théorie de construction d&apos;un simulateur elastoplastic d’hydrofrac à trois dimensions entièrement couplé, en particulier conçu pour les roches non consolidés, est présentée dans ce papier. Ce nouveau modèle calcule les dimensions et les changements dans la forme de(s) fracture(s) étant créé pendant les opérations de fracturation hydraulique. Aucune supposition n’a été faites en ce que concernant la forme ou la géométrie

Jan 1, 2003

By M. A. de Ruijter

Bracken Gold Mine ceased mining operations during February 1993. The clean-up of the gold plant, which had already commenced in September 1992, was completed in January 1994. The mine had been in production since early in 1962 but, because of a decreasing ore reserve, increasing working costs, and a relatively low gold price, closure had become inevitable. The closure had been under consideration since August 1983, when the first estimate was made regarding the amount of gold that could be recovered from the clean-up operations. Until that stage, it had been accepted that Bracken would conform to the commonly accepted idea that 0.25% of the total gold produced by the plant would be recovered during the clean-up operation. This idea was dispelled when the initial estimates indicated that the gold recovery would vary between 0.049 and 0.095%. The actual figure realized was 0.036%, showing that there is no generalforl1ula for the calculation of the amount of gold to be recovered from such clean-up operations. During the clean-up operation, Bracken became an unexpected contributor to the State&apos;s Reconstruction and Development Programme by having al the mine and plant buildings purchased and converted into a community development centre. This shortene4 the time available for the clean-up, and several of the buildings remained intact. However, the foundations and sub-soils had been sampled extensively, and it is considered that no recoverable gold was left behind.

Jan 1, 1996

By S. Erçayhan, B. Derin, O. Yücel

In the present study, the production of low carbon ferrochromium (LC-FeCr) from chromite concentrates containing 55.09% Cr2O3 and 13.06% Fe via metallothermic process was investigated. For this aim, in the experiments, the aluminothermic reduction method was applied in an open crucible. 10% CrO3 + 10% NaClO3, 15% CrO3 + 5% NaClO3 and 20% CrO3 by weight of concentrate were added to the charge, respectively and the charge was reduced with aluminium. The experiments were done with the additions of 90, 100, 110, 120 and 130% of stoichiometric requirement of aluminium. The highest Cr content in alloys was found as 71.7% at 120% of stoichiometric Al whereas the highest Cr recovery in alloys was found as 71.8% at 130% of stoichiometric Al when the both charge contained 15% CrO3 + 5% NaClO3 by weight of chromite concentrate.

Jan 1, 2004