Search Documents

By Karl-H. Daum

Over many decades, numerous feasibility studies have demonstrated, that the production of sulphuric acid remains the most viable option of sulphur recovery from smelter off gas and abatement of SO2 emissions to the atmosphere. This is particularly more pronounced as smaller but more concentrated off-gas flows are to be treated from smelters, and enhanced sulphuric acid processes be-come available. Off-gas handling systems represent a significant capital and operating cost burden to the metallurgical operation. Modern pyrometallurgical smelter processes for sulphide ores based on the use of oxygen-enriched air, produce relatively small off-gas flows with high SO2 concentrations in the smelter gas of 30?60 %-vol. of SO2. This is a prerequisite for substantial cost reductions in the smelter off-gas handling and treatment system. As an alternative to sulphuric acid production, numerous scrubbing concepts with alkali or dual-alkali combinations as well as organic absorbents have been proposed. Also the reduction of the SO2 to elemental sulphur has frequently been studied. Very few of those alternative processes have been built in industrial scale, but all were generally characterized by none sustainable operation due to cost reasons, problems with issues related to chemicals used and by-products or poor availability. Thus the traditional concept of converting the SO2 to sulphuric acid is most common, although regarded as un-economic, but is at least a proven, environmentally sustainable and reliable way of sulphur gas processing. With the high acid price levels these days, operating companies are even able to generate significant revenue with their otherwise ?fatal? acid. Even though smelter gas is available at high SO2 concentrations, in a conventional acid plant one would add large amounts of air to dilute the gas down to a suitable concentration of 12?13 (14) %-vol. SO2. Dealing with higher SO2 concentrations in smelter acid plants is not feasible, as the gas exit temperature of the catalytic oxidation step would exceed the allowable limit for the catalyst. Hence those conventional acid plants are characterised by the use of large equipment, i.e. high capital cost, high energy consumption and limited flexibility. The newly developed LUREC® process can handle off-gas with significantly higher gas concentrations, even in excess of 25 %-vol. SO2. It is entirely based on well-proven equipment and unit operations. The first industrial application, operating with 16?18 %-vol. SO2 will be presented in this paper, along with the fundamentals of the process. It will be demonstrated that the process requires inherently lower capital cost, fewer operating cost, offers better energy recovery and lower emissions as compared to conventional design. While most existing smelter acid plants would have some built-in spare capacity, any significant increase of smelter capacity, say of 30%, can basically not be accommodated without installing an additional parallel new acid plant unit. An add-on LUREC® module can remove any such restriction, while simultaneously de-bottlenecking the existing acid plant. The paper also discusses the present status of the add-on technology for smelter acid plant expansions with respect to technical boundary conditions of different process alternatives as well as economic and environmental aspects. Application of the LUREC® process to modern smelter acid plants operating with high SO2 gas concentration, will lead to a substantial reduction of the specific size of the gas processing equipment and equivalent savings in capital and operating cost. The LUREC® process is patented worldwide by OUTOTEC.

Jan 1, 2009

By FJ Fourie

Lack of success in attempts to develop viable large-scale continuous mining techniques (previously reported by MP o?Hare in 2004) led the team at Kopanang to explore alternative solutions to their ongoing need for improved profitability. By calling into question the basic assumptions around the stoping of narrow tabular deposits of gold, fragmentation and more particularly, the range of sizes of blasted rocks eventually became the focus of serious study. This background is briefly covered before the sequence of the investigations is described. So far, some of the ?rules of thumb? derived over the years have been confirmed, albeit for totally different reasons than those from which they arose; some persistent theories on gold loss have been debunked, and some new ones have been developed. Blast optimisation (the means by which desired fragmentation is achieved) has been based on tailoring energy to rock characteristics in an innovative packaging design for explosive charges. Other beneficial consequences arising from the work include increased production volume and gold output, a decrease in milling consumables, and the Mine Call Factor is being maintained at a level previously thought to be unachievable. This is very much a ?work in progress?, and it is empirical in its approach. The authors are the first to recognise that they have yet to determine in absolute measures why some of the innovations have had the effect that has been observed. Also that there other factors which may contribute to the result, but which have not yet been examined. Nevertheless, publication of the work completed so far is felt to be justified; firstly on the basis of the achievements themselves, and secondly for the benefit of others in the industry for whom similar problems exist. Statistical reports of the effects on the business are included at the end of the report. Detailed laboratory test results have been separated into an appendix, leaving salient graph and tabulated data in the main text. For ease of reading, units of measurement for particle sizes below 1mm are described in microns, but millimetres have been retained for those at that size or greater.

Jan 1, 2007

By C. van Buuren, R. F. Sandenbergh, P. J. Pieterse

Gold extraction from double refractory ores using conventional cyanidation has proved challenging due to the presence of carbon within the ore. Carbon particles may adsorb leached gold resulting in gold losses to tailings, i.e. preg-robbing the gold. Some of the preg-robbed gold may be recovered by the application of the HiTeCC process on CIL tailings. The process developed by Metso Outotec involves caustic conditioning of the tailings followed by high temperature desorption and simultaneous adsorption onto virgin activated carbon of the aurodicyanide. The aim of this paper was to develop a model for the HiTeCC process which could be used to support design and optimisation of the process, based on the adsorption characteristics of the preg-robbing and activated carbons. It was found that accurate experimental isotherms can be used to predict the behaviour of preg-robbing adsorbents. Comparing the sorption isotherms for preg-robbing ores and activated carbon indicated similar dependencies on temperature and ionic composition. Applying boundary layer diffusion models coupled with experimentally determined isotherms for preg-robbing ore and activated carbon, the HiTeCC process was successfully simulated at experimental and plant scale. Gold recovery of the HiTeCC process is most sensitive to the temperature of desorption and gold loading of the activated carbon, with increased temperature improving recovery. The gold loading of the activated carbon in the process should always be maintained lower than that of the normalised pregrobbing carbon gold loading to ensure favourable conditions for adsorption onto the activated carbon. Caustic conditioning stabilises adsorption and is therefore not necessary during the desorption stages of the process, but potentially improves recovery during adsorption. Keywords: Preg-robbing, refractory gold ores, HiTeCC, aurodicyanide

Jan 1, 2020

By S. T. L. Harrison, M. A. Fagan-Endres, M. D. Odidi

This paper investigated the ability of eight common hydrodynamic models – tanks in series (TIS); three compartmental model configurations (Comp-1, Comp-2 and Comp-3); advection dispersion (AD); piston exchange (PE); piston exchange-diffusion variant (PE-D) and piston dispersion exchange (PDE) – to simulate residence time distribution (RTD) data generated from tracer tests conducted on a packed bed composed of different size fractions of malachite ore. The chosen materials were packed into 0.1 m internal diameter columns to a depth of 0.155 m and drip irrigated at 5.11 L/ m2.h. The four size fractions tested were: 1 – 2 mm, 2 – 2.8 mm, 4.75 – 5.6 mm and 13.6 – 16 mm. Step-up tracer tests were conducted using potassium chloride. The results showed that an increase in the particle size of the malachite ore had little effect on overall bed porosity but resulted in significantly lower values of liquid holdup. Two of the monoporosity models tested, TIS and Comp-2, were the most affected by the changes in bed size fraction due to their inherent inability to cope with high degrees of non-ideal flow present in the larger particle systems. The dual porosity models performed relatively better with the dominant trend being a reduction in dynamic to total liquid holdup as particle size increased. Of all the models tested, the PDE and TIS models were the best and worst performers respectively, with R2 values averaging at 0.997 and 0.679 across all experiments respectively. Though the Comp-1, Comp-3 and PE-D models also performed well, the addition of the dispersion coefficient to take the axial spread of fluid molecules into account added a layer of complexity that allowed the PDE model to perform relatively better than the other dual porosity models. Keywords: Heap leaching, packing material, hydrodynamics, malachite, PSD, modelling

Jan 1, 2020

By A. Ndibalema

Blasting is one of the primary functions in any mining operation. Despite the fact that it generally constitutes between 30% to 40% of the mining cost, blasting also affects the cost-effectiveness of down steam activities such as load and haul, safety, tyre life for haul trucks, crushing and milling. Over the past three decades, significant progress has been made in the development of new technology in an attempt to reduce costs and increase efficiencies and productivities of blasting activities. Mine to mill project concepts have paved the way for such developments. Other advancements have included sophisticated computer modelling technologies for blast design and performance analysis. There are many factors that affect the cost of blasting. These can be divided into two categories; controllable and uncontrollable. The latter includes costs related to depreciation, exchange rates and government legislation with the former category including contractor versus owner blasting costs, efficiency and productivity issues and product quality. Controllable factors have one common denominator which can determine their outcome; direct site management influence. Excavator digging rates and final floor conditions give a good measure of blasting performance, and alternatively cost effectiveness. This paper is aimed at benchmarking against a blast optimisation database, understanding the relationship between costs and identifying opportunities for improvement. Understanding the cost drivers allowed a typical cost model to be developed for the current mining operation at Geita and this was used to drive continuous improvement with emphasis on closing the information loop.

Jan 1, 2008

By J. Abdollahisharif

Optimal cut-off grade is one of the most critical parameters in open-pit mine design because it defines the ore and waste and thus determines the maximum net present value possible from the mining operation. Although the algorithm presented by Lane (Choosing the optimum cut-off grade. Colorado school of Mines Quarterly, vol. 59, no. 4, 1964. pp. 811-829) is the most common one in the literature, it does assume constant capacities of the mine, processing, and refinery. In this paper, Lane's original algorithm has been modified to consider variable processing capacities in order to determine the optimal cut-off grade in open-pit mines. The new algorithm proposed here is compared to both Lane's original algorithm and to a previous modification that incorporated rehabilitation costs into the cut-off grade estimation. The algorithm proposed here that permits variable processing capacities is shown to be superior for the determination of optimal cut-off grade to both these previous versions. In addition, a computer-based program was developed in Microsoft Office Excel© to calculate the optimal cut-off grade as proposed here.

Jan 1, 2012

By R. M. Kear

The Mine Planning Department of Rio Tinto Management Services South Africa has installed an interactive graphic work station to assist with the long- term mine planning of Palabora. Previously, the planning was drafted manually after which digitised areas were fed to the computerised ore reserve system. This method was slow, tedious, prone to error and was becoming incapable of meeting new planning parameters. To assist the planning engineer an interactive graphic system was developed, and subsequently a rewrite of the ore reserve reporting system was incorporated. The work station has reasonable computing power and, when not utilised for graphics, has been found suitable for the evaluation of various non- graphical applications, a recent example of which was the in-pit crusher investigation. This project, in contrast to similar studies, was not based on a fixed system throughput but rather on a well designed system which would be operated at its own optimum capacity. In order to determine the probable throughput a simple simulator was designed to model the crushing and conveying system. This paper describes the investigation into the specifications for the work station, the selection of hardware, operating system and languages. An over- view is then given of the graphic planning system. The in-pit crushing simulation has been included as an example of how the work station has been applied to other, non graphical, applications.

Jan 1, 1987

By A. Karadogan

Decision making can be defined as the selection process of the best routes from the alternatives in order to achieve a goal, and mostly involves uncertainty. Additionally, mostly qualitative variables (weak rock, massive ore deposit, etc.) are in question. Mine planning engineers often use their intuition and experience indecision making. As fuzzy set theory has been used since 1970, these uncertainties are easily evaluated in the decision-making process. Real world study is decision making under vague constraints of different importance, involving uncertain data (qualitative variables), where compromises between antagonistic criteria are allowed. This paper presents a new approach in the selection of an underground mining method based on fuzzy set theory for the Ciftalan Lignite Site located close to Istanbul in Turkey. The physical parameters such as geology and the geotechnical properties of ore, hanging and foot wall, economic effects, environmental effects, are established with field and laboratory tests together with the determination of other qualitative variables. Meanwhile, some qualitative variables dealing with the matter was described according to the view of a number of experts. Then fuzzy set theory is applied to these parameters, considering the available underground methods in order to choose the proper method. At the end of the evaluations, the room and pillar method with filling was determined as the most suitable method for the test site. Keywords: Decision making, fuzzy set theory, underground, mining, method selection

Jan 1, 2008

By D. G. Hulbert, I. K. Craig, S. P. Moult

This paper describes the application of a throughput-optimization strategy to a run-of-mine (ROM) mining circuit that is controlled by a multivariable controller. The objective of the optimization strategy is to maximize the ROM throughput at a specific product size. Results from long-term tests, which indicated how 'steady-state' optimization can be achieved, were used for the throughput optimization. This strategy increased the ROM throughput by 11,4 per cent. A mathematical relationship relating residue to grind can be used to calculate the economic benefits of stabilizing control. It was estimated that the throughput optimizer could result in an increase in gold production worth Rl,4 million per annum when waste rock is treated, and that stabilizing control could increase the treatment plants annual revenue by R150 000. If the extra milling capacity brought about by the optimizer is interpreted as a capital cost saving of 10 per cent on the price of a new mil, then a capital cost saving of RI million could result.

Jan 1, 1992

By K. Eckert, A. E. Sommer, S. Heitkam

One highly relevant challenge in flotation is the recovery of fine particles. Due to their low inertia, these particles are mostly pushed aside by rising bubbles. Consequently, they typically exhibit low probability for bubble-particle collision and thus, a poor recovery rate. In this work, the trajectories of fine particles in the vicinity of rising bubbles were investigated. The measurements considered a chain of millimetric bubbles in a rectangular container filled with deionized water. The flow field near the bubbles was measured by tomographic particle image velocimetry (TPIV), employing fluorescent tracer particles of 33μm diameter. Subsequently, trajectories of larger particles and bubble-particle collision events are recorded with 4D particle tracking velocimetry using a high temporal and spatial resolution. The results reproduce the well-known collision of particles on the leading edge of a rising bubble. Additionally, collisions on the tailing edge were observed in cases with a low Stokes number. The TPIV results demonstrate, that the high turbulent kinetic energy in the bubble wake allows particles to divert from the fluid streamlines and collide with the tailing edge of the bubble. The tailing edge collision probability increases with the bubble Reynolds number and decreased particle inertia. Overall, the investigation shows that the collision of fine particles in the bubble wake should be considered for the development of further collision probability models. Furthermore, the importance of turbulence on the fine particle flotation was demonstrated. Keywords: 4D Particle tracking velocimetry (PTV), tomographic particle image velocimetry (TPIV), fine particle flotation, collision probability, bubble-particle interaction, wake induced collision

Jan 1, 2020

By A. J. S. Benade, N. B. Strydom, W. H. Van Der Walt

The performance of the improved microclimate suit by N. B. STRYDOM M.Sc., Ph.D. (Illinois) (Fellow), A. J. S. BENADE, B.Sc. (Potch.) (Visitor), and W. H. VAN DERWALT, M.Sc. (Potch.) (Visitor) An assessment has been made of a new improved microclimate jacket when worn by men working for four hours in a hot environment (wet-bulb temperature of 33,9°C and an air speed of 0,4 m/s) and at work rates of 35 Wand 70W. At the lower rate of work in heat, the use of microclimate suits maintained the rectal temperatures of the men at almost the same level as when they worked without that protection at normal room temperature (20°C). At the greater rate of work without protection, the men's rectal temperatures rose to 40°C within two to three hours, so that they could not complete four hours of work. When they wore microclimate suits, they easily completed four hours of work, and their rectal temperatures were lower than when they worked at the lower rate without protection.

Jan 8, 1975

By M. C. van Blerck

Since its introduction in 1956, the capital allowance has been increased and its scope widened. At present the allowance is 10% in respect of two categories of gold mine (post-1973 and other deep level mines) and 6% in respect of natural oil mines. The allowance is a significant incentive encouraging the development of new mines, to the benefit of both the fiscus and the South African economy as a whole. A strong case can be made for increasing the amount and the scope of the allowance. The Margo Report recommendation that the allowance should possibly be replaced with a system of direct grants is, in theory, neutral. However, practical difficulties exist with cash grants and there is concern with the implications elsewhere in the Margo Report that the capital allowance is excessive and that, as a consequence, such direct grants will represent an effective reduction in this incentive. I believe that the best course of action would be the retention (and enhancement) of the capital allowance and the expansion of its scope.

Jan 1, 1988

By B. Kweto, W. Legodi, C. van Buuren, R. F. Sandenbergh

Biooxidation, and especially the BIOX® process, is a viable, economical and environmentally attractive process to liberate gold hosted in refractory sulfides by oxidatively leaching the sulfides as soluble sulfates. The moderate to high cyanide consumption during the subsequent cyanidiation of the washed biooxidation residue is probably caused by the retention on the solids of labile sulfur species formed during the multi-electron oxidation process of sulfide to sulfate. In this work the nature and mode of transport of the cyanicides in the products from the mesophile BIOX® process, as well as that of the combined mesophile and thermophile bioxidation process, were considered and strategies to reduce cyanide consumption evaluated. It was found that labile sulfur species were the major cyanide consumers in the mesophile biooxidation product, forming predominantly thiocyanate during cyanidation. The transfer of labile sulfur species was only marginally decreased by more intensive and thorough washing of the biooxidation products with warm and hot water, intense stirring and warm caustic leaching. A major reduction in cyanide consumption was however achieved by using a combined mesophile and thermophile biooxidation process. Keywords: Biooxidation, refractory sulfide gold ores, cyanide consumption

Jan 1, 2020

The first commercial Corex Plant in the world capable of producing liquid iron from a variety of low-grade coals was commissioned in 1990 by Iscor at its Pretoria Works. The Plant has an annual production capacity of 300 kt of liquid iron. Because the compulsory use of coke is eliminated, the new method offers an economic and revolutionary alternative to liquid-iron production. Coke used in traditional blastfurnaces is manufactured from high-grade coal of which the available sources in South Africa are limited.

Jan 1, 1991

By T. R. Twiddle

The sequential flotation of copper, lead, and zinc in that order is practised at the Broken Hill Concentrator of the Black Mountain Mineral Development Company (Pty) Limited. The minerals of economic importance are chalcopyrite, galena, and sphalerite, with associated silver. The targets for recovery and grade in the lead and zinc circuits were reached quite readily after start-up, but the copper circuit failed to meet the performance requirements with respect to both recovery' and grade. The original circuit for copper is described, together with the subsequent investigations and modifications that were required to produce acceptable metallurgical results. Circuits for semi-bulk and bulk flotation were tried, and the effect of the oxidation state of the pulp was investigated extensively and found to be all important. Tests were conducted on the depression of lead in the cleaning circuit by the use of sodium dichromate, and on heating with the addition of sulphurous acid. The latter technique was ultimately the more successful. The proportion of fresh to recycle water entering the circuit was shown to be important. The original reagent regime was altered and augmented. A more recent problem has been the contamination of the ore with cemented backfill. The changes that this new condition necessitated are described. Success has been measured by a continuous improvement in the economics of the copper recovery.

Jan 1, 1984

Dr Hanekom I would like to thank Dr Hanekom for a very valuable -contribution; it summarizes careful statistical test work which was an omission from the original paper. Minor manipulative details, the application of a finer measuring graticule and the reproducibility of the procedure are described and the author accepts with gratitude this ready-made answer to the first and second points of Dr Finkelstein's comments. Mr Williamson The author is indebted to Mr Williamson for historical background to early experiments with captive bubbles and it is conceded that the original paper described another avenue to an old flotation goal. The field of application is restricted and we would not work with all the material illustrated by Klassen and Mokrousov.1 The particles in Figure 74(a) are far too large and irregular; material in Figures (b) through (d) could be employed for semi-quantitative work in spite of obvious disadvantages; Figure 74(f) represents ideal conditions. We have studied this mineral-collector association in the size range given and results were satisfactory. For larger pick-up values than that illustrated an induction period of more than 0.01 second was required (actually 90 seconds for a maximum bubble load at pH 7.2 with 25 g/t oleic acid!). REFERENCE 1. KLASSEN,V. I. ANDMOKROUSOV,V. A., 'An Introduction to the Theory of Flotation,' Butterworths, London, 128, Figure 74, (1963). Mr Slander and Mr Kooij Researchworkers in a practical mineral development laboratory may well ask what contribution yet another micro-technique can make to the problems of large scale ore flotation. The obvious answer is that there can be little or no direct translation of results obtained with a static system comprising air bubble and pure mineral concentrate to a dynamic multicomponent operation. However, for the rapid study of one mineral or variable some small scale apparatus must be employed. In the Anglo American Research Laboratory, the described semi-quantitative pick-up technique has filled a niche and requirement. For ourselves, the system represents a definite advance over the contact angle approach in that samples of minerals conventionally prepared in laboratory or plant may be used, which is a desired step nearer flotation. We bracket this pick-up apparatus with the Hallimond tube1,2 and evidence is now to hand that pick-up curves may often be related to flotation recoveries obtained from this tube with pure mineral samples and from the Fuerstenau glass cell3 with the total ore at low pulp densities. We would not wish to push the comparison of pick-up results past this stage for there is much published development work relating the Hallimond and Fuerstenau equipment to flotation on a larger scale. At the Anglo American Research Laboratory for example, a test in a Fuerstenau cell is often the final integrated step in laboratory investigations into the TORCO process. As these contributors rightly point out, pick-up tests can offer nothing towards the determination of optimum reagent dosage for large scale flotation. This is the function of bench-type flotation cells with up to 2 kg batches of ore. At this level, collector performance and consumption and physical properties of the froth are more comparable with large scale operation. The pick-up cell, Hallimond tube and Fuerstenau cell are essentially equipment from which the effect of control parameters can be derived. For the assessment of collector behaviour, the pick-up approach is rapid and economic and we offer here a pick-up contribution towards the flotation of Witwatersrand uraninite with anionic sulphonate collectors (Fig. 1). A large crystal of the local mineral was not available for contact angle measurements and owing to its low concentration in Witwatersrand gold-bearing reef, a small sample was laboriously obtained by tabling, superpanning and heavy medium separation. The pick-up values of a sized portion were determined by the described procedure in the presence of nine petroleum sulphonate collectors. Responses fell into two categories; "normal" curves (dotted lines) and a group of collectors which were very active at low concentration but showed depression of pick-up at the higher levels. The similarity in form of curves for collectors 1 and 3 was apparent and reagent 3 was found to be simply a diluent of 1. Otherwise, collector performance could not be related to either sulphonate content or molecular weight of the active constituent. However, reagent 2 gave the best performance in a bench flotation cell with regard to overall uranium distribution but, as predicted, reagent 1 yielded the best grades. Later pick-up tests with a similar class of collector have suggested that depression at high reagent concentrations may be due to "bubble armouring" whereby an hydrophilic air-liquid interface is produced which rejects mineral attachment. Prolonged investigations with the same mineral sample have not been required except for the study of uraninite which lasted for more than a year. Variation in pick-up response could not be detected over this period during which the mineral was stored under distilled water. The only visible change was a small efflorescence of lead sulphate, readily removed and decanted off after stirring with a glass rod in the sample tube. In its main role as a "trouble-shooter", the pick-up method is applied to fresh mineral samples taken from a flotation plant and soon discarded or from a current run in the ore-dressing laboratory. After production of a concentrate, a solvent cycle, as described in the original paper, may be required to remove heavy organics or other contaminants as when a sample has been taken from circuits containing detergents, grease or return flotation solutions. This solvent cycle will not remove chemisorbed collectors and we prefer not to work with minerals from flotation tailings or with pick-up cell discards. It has been thought preferable to answer the contributors' other apprehensions with regard to sample preparation, activators, depressants and the physical properties of reagents in another paper on offer to this Journal. It covers pick-up investigations relating to the flotation of nickeliferous pyrrhotite at two Anglo American Corporation Mines, which approach seemed the correct orientation for a reply to another contributor to this paper.

Mr R.D. Beck attended Felsted School in Essex (England) and the Royal School of Mines, University of London, where he ob¬tained the degree of B.Sc.(Eng.), A.R.S.M. in mineral technology in 1969. His other qualifications include Pr.Eng. (1976), C.Eng (1975), and M.I.M.M. (1975). He joined Gold Fields of South Africa in December 1969, and worked at Doornfontein Gold Mine, Greenside Colliery, West Driefontein Gold Mine, Zwartkloof Fluorspar, Gold Fields Laboratories, Gold Fields Head Office, Libanon Gold Mine, and Kloof Gold Mine before being appointed Plant Superintendent at East Driefontein Gold Mine in January 1973. In June 1976, he was transferred to Head Office as Group Metallurgist, followed, in November 1978, by his secondment to Black Mountain Mineral Development Company as Plant Superintendent, to take on-site responsibilities for staff training, plant commissioning, and start up. On his return to Head Office in July 1980 as Assistant Consulting Metallurgist, he was responsible for Black Mountain and new metallurgical projects. Then, in January 1984, he was appointed General Manager of the Zinc Corporation of South Africa Limited. He became a member of The South African Institute of Mining and Metallurgy in 1973.

Jan 1, 1985

Contribution by W.D. Ortlepp? The authors are to be commended for publishing their interesting view of an aspect of deep-level mining technology that is becoming increasingly topical and vitally important. This is the second paper in the journal within four months to deal with the crucial needs of highly stressed tunnel construction, the other being the paper on the design and testing of rockburst support in the February 1994 issue. Both papers re-examine important innovative ideas that are more than a quarter of a century old, but that have not yet been implemented by the industry. It is to be hoped that they will now, taken together, serve as a new starting point for a definitive examination of what constitutes proper tunnel-support design. There can be no question that the need to develop a design approach that will solve the problems of tunnelling at depths of 4000 m or more has now become crucial. Contribution by A.M.J. Vervoort* Although the overall paper is very interesting and the future research (comparison between numerical simulations and underground measurements) will be very challenging, I would like to make some remarks regarding the set-up of the numerical model and the interpretation of the elastic results.

Jan 1, 1995

By W. P. Channon, J. B. See

The reduction of fluxed and non-fluxed manganese, ores by ferromanganese silicide by W. P. CHANNON, B.Se. (Eng.), M.Sce. (Eng.) Wits. (Visitor) and J. B. SEE, B.Sce. Hons., Ph.D. (New South Wales), M.I.M., M.I.M.M. (Member) In South Africa, medium-carbonferromanganese is produced by the silicothermic reduction of slags containing MnO. Because of its importance in the production of refined ferromanganese, a study was made of the reaction between a premelted South African Mamatwan manganese ore and various ferromanganese silicide reducing alloys, with and without fluxing agents. Because the reaction is highly exothermic, temperature appeared to have little effect on the rates of manganese transfer to the metal and silicon transfer to the slag, and an apparent equilibrium was rapidly approached. Limits were determined for the favourable effects of fluxing additions, and manganese recoveries were found to decline as the ratio of CaO to MgO decreased as a result of the addition of basic flux. The recovery of manganese passed through a maximum, and subsequently decreased for variations in the ratio of alloy to ore. The results obtained, when used in conjunction with other physicochemical data related to this slag-metal system, suggest optimum slag compositions for the efficient manufacture of medium-carbon ferromanganese from Mamatwan ore.

Jan 3, 1977

By S. C. Essack

"Tenova Pyromet has designed a slag flow control valve that throttles the flow of slag in order to control the tapping rate. In addition, the valve has proved to be a reliable tool for closing the furnace during normal operation as well as emergency conditions. The valve makes use of a water-cooled copper element that is hydraulically operated and attached directly to the slag tap-hole. The equipment is compact, simple to maintain, and cost-effective to purchase and install. It can be retrofitted to most existing slag tapholes with minimal modifications.During the development phase, finite element method (FEM) and computational fluid dynamics (CFD) software was extensively used to model predicted operating conditions and to optimize the design. By using CFD, it was possible to understand the flow patterns inside the copper and to optimize these to improve overall performance.The slag valve was installed and commissioned in 2011/2012 on a 12 MW furnace as part of a new mineral wool plant in the USA. It has performed well since its installation despite operating under conditions in excess of the original design criteria. During the first 6 months of operation it quickly became the preferred method of closing the furnace, although the design intent was for occasional use during emergency shutdown conditions only. Since replacing the mudgun as the main method of closing the tap-hole, a few minor changes have been made to the installation in order to cope with the increased heat load and much higher frequency of use."

Jan 1, 2014