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By G. A. Wiebols, N. G. W. Cook, N. C. Joughin

Discussion R. E. Rarnes (Member): The original concepts and the pioneering work now brought to the practical test stage by the Mining Research Laboratory team deserve our highest praise. The authors of this comprehensive paper rightly stress the urgency of establishing the extent to which the apparent potential can profitably be realised in practice. It is to be hoped that adequate funds will be made available by individual mining companies, the Chamber of Mines and manufacturers to attract the necessary staff and maintain the high rate of achievement of the last two years. From the Seventy-Eighth Annual Report of the Chamber of Mines and its members we see that, in 1967, with a Working Revenue of R759.8 million from gold and R54.6 million pit mouth coal sales and with profits from gold and uranium and pyrite of R307.9 million, only R1.9 million was spent by the Chamber on all forms of Research. It is considered unlikely that the associated mining companies and manufacturers exceeded this investment expenditure. Assuming a total of R4 million spent by the industry on Research and Development, this is less than half of 1 per cent of sales of gold and coal. This percentage, so low in comparison to North America and Europe, is no worse than that of Exploration expenditure which, in 1967, with South Africa's total mineral production of R1,287 million, was estimated to have been R6 million (Pretorius 1968). In a primary industry with ever present depletion of ore deposits and with cost escalation, expenditure on Exploration and on Research and Development is not a risky luxury but a tactical obligation. The potential rate of return on research expenditure into rock breaking is high. Stores consumed by gold and coal mines, members of the Chamber of Mines, totalled R316.2 million in 1967. Except for purchased power costing R42.2 million the highest cost group was explosives, drills and drill steel totalling R33.8 million or 10.7 per cent of the total stores consumed. In the paper under discussion it is claimed that the low 'effective stoping width' should greatly reduce the likelihood of rock falls or rock bursts to the extent that permanent support can be dispensed with. Insofar as this narrow cut is only 12 in. in advance of the working area which, with a 10 in. channel, is unlikely to be much reduced in width from that achieved by current methods, this claim is not readily understandable. Were it to have been based upon the regional support gained from packed waste it would have been more acceptable. Pre-developed stope drives may give serious trouble at depth and for this reason it is questioned whether a stoping area can avoid periodic sub-development blasting-the spoil and fumes from which will interfere with the rock flow and continuous mining of the rock-cutter. If, in the mining method proposed by the authors, stope drives are cut as small as possible (6 ft by 6 ft) then 25 per cent of the total tons handled (excluding resued waste) and 5 per cent of the gold will be blasted conventionally in the stope. At this stage one cannot envisage tunnelling machines economically or practically capable of such work. The various methods described by the authors and subsequent contributors for breaking waste are most interesting. It was noted that the 'bull wedge' and 'explosives' in Fig. 1 of the paper were no further from the ideal point 'A' than was 'cutting'. The writer considers that the bold and imaginative steps taken by the Mining Research Laboratory Team, the mining companies and the manufacturers concerned will eventually lead to a successful rock cutting machine with universal application largely independent of rock type. This may take many years. In the meantime other methods of improving productivity of saleable metal by rock breaking teams should be investigated even if such methods have local applications only. In 1955 the writer conducted tests with a wire saw similar to those used in quarries in the Northern Transvaal and elsewhere. Jeppestown shale, the immediate footwall of much of the East Rand gold field, was cut at the rate of 6 in. per hour using sand, water and a special endless rope driven by a low h.p. motor. A hypothesis on its application was submitted to the Office of the Government Mining Engineer in 1955 and to other mining institutions in 1966 after the writer returned to South Africa. By inference, rope sawing was classed as less promising than other methods tested in the Orange Free State Goldfields (Parker 1969). With highly resilicified hanging and footwall quartzites this was not surprising and confirmed the writers findings when testing hanging wall quartzite from the East Rand in 1955. The relatively uniform conditions, the low strength, hardness, and silica content of the Merensky Reef platinum deposits (Gray and von Bardeleben 1969) and in particular, the existence of overlying Merensky pyroxenite (Cousins 1964) make this and the East Rand attractive areas for larger scale testing of wire saws. It is envisaged that in suitable rock types the 5/8 in. slot would be advanced down dip or down a minor dip. In undisturbed areas 'faces' of up to 200 ft in length could be cut several feet in advance of breaking which could then consist of light blasting to the second free face or some of the methods now being tested for breaking waste in rock-cutting operations. A wire saw is an inexpensive and simple machine which, in some areas, could make significant and early gains in rock breaking efficiency as well as in ground and stoping width control.

Jan 5, 1968

By J. Baklík, V. Rak, W. Kolarcík, Z. Franc

The paper deals with hydraulic transport of thin and dense slurries. Based on mathematical modelling of three-phase flow, dynamic properties of the systems for pumping of thin slurry are solved. Theoretical calculations were experimentally verified both at the testing hall and in real operation. For thin slurry (e.g. bauxite slurry), a three-chamber hydrohoist with float is used. Dense slurry (e.g. ash and water) is transported from the power plant to a deposit site by means of a three-chamber hydrohoist with sealed separator (piston).

Jan 1, 2007

By H. Wagner

The discovery of the gold-rich Main Reef Group of reefs on the farm Langlaagte in 1886 marked the opening of the richest and most extensive goldfield in history. Mining of the narrow reefs progressed from the outcrops to more than 3,5 km, and has made South Africa the foremost mining country in the world. Changes in the political and socio-economic climate create a new challenge to the gold-mining industry. The most significant of these is the emergence of Black trade unions, with their demands for higher wages, improved safety, job advancement, and abolition of discriminatory legislation. Industrial conflicts on gold mines have risen significantly in the past ten years, and are likely to continue in the future, although the causes of these conflicts may change. Events largely outside the control of the gold-mining industry, such as the improved dollar price for gold and the depreciation of the value of the rand relative to major currencies, have enabled the industry to meet the demands for higher wages and to narrow the wage gap. However, labour productivities have improved insignificantly and are a cause of major concern for the future development of the industry. Continued pressure in the industrial-relations area, the inflation of working costs, and the extension of mining operations to greater depths and more difficult areas create major challenges to mine managements and mining engineers. Significant improvements in mining technology, environmental control, and safety will have to be made to ensure the long-term survival of the industry. The unavoidable introduction in deep gold mines of mechanization on a larger scale to meet the requirements of higher labour productivities and a better and safer underground environment will have a major impact on the cost and management structure of these mines. Engineering will assume a more important role on deep gold mil1es, and management styles and structures will have to be adapted to accommodate these changes. The supply of suitably qualified engineers, technicians, and artisans will be a crucial factor controlling the rate of technological change. The industry is well-geared through its Group system to meet these challenges, and to benefit from the experience gained in the operation of highly mechanized mines. However, changes take time to implement and there is very little time left.

Jan 1, 1986

By J. P. Coetzee

Dr. K. Gebhard In his excellent and comprehensive survey on the future development of the Iron and Steel Industry in South Africa, Mr. Coetzee has also mentioned the direct reduction processes as a possible method of converting iron ore to iron. Mr. Coetzee has made it quite clear that at this stage one should not consider these processes as a 'magic formula' which herald a 'break through'. I agree entirely with this point of view. Once upon a time I held the opposite view, but in the meantime I have revised my views. To do so, I had 40 years at my disposal. In 1930 I had the privilege of being connected with the first commercial sponge iron plant in Germany1 - a plant which was capable of producing 40 tons of sponge iron a day. We were determined to prove that the blast furnace would become obsolete in the near future. We considered it as an unnecessary detour to convert iron ore first into an iron-carbon alloy, which contained further quite a number of undesirable elements, and to remove this carbon by an endothermic process. Direct iron-making seemed to be a logical short cut. Unfortunately, we were not able to establish this 'short cut,', but the life of our sponge iron plant was cut short. We had solved all technical but not the economic problems. Basically the positon is still the same today - mainly economic and to a lesser degree technological aspects determine the rate at which sponge iron plants are installed. More has been said and written than done about direct reduction plants. I think that there are two main reasons for this stagnation in the introduction of direct reduction processes. The first reason is that modern blast furnaces are just too good and too economic ever to be replaced by direct reduction processes on a large scale. This, of course, already limits the field of application for direct reduction plants. The second reason is that sponge iron production and utilisation is based on technical considerations, which increase not only the production cost of sponge iron but also the conversion cost to steel to such an extent that under special conditions only are direct reduction plants an economic proposition. Are these technical considerations still as forceful and valid today as they were when 'direct steelmaking' was first contemplated? The development of sponge iron production as a method of direct

Jan 3, 1971

By M. Moys

The use of temperature measurements for unusual applications in the analysis and control of column flotation and milling is discussed. This is followed by a review of some of the authors current research in the application of the discrete element method to the modelling of particulate processes, specifically milling, fluidized beds and chute flow.

Jan 1, 2005

By J. -M. Milazzo, T. Wallmach, S. B. Blancher, A. Amalric

In a FeNi electric arc furnace, the physical properties of the slag (chemistry, viscosity, liquid/solid/gas ratio) are of crucial importance in controlling the efficiency of the process. In this paper, we studied the slag crystallised on the walls of a semi-industrial furnace (freeze-lining), the slag rising through the burden during foaming phenomena, and the casted slag. The freeze-lining is the layer of oxides that crystallises on the refractory wall of a furnace, protecting it from chemical and mechanical aggression. It records changes in furnace conditions (slag chemistry, temperature, oxygen fugacity), which makes it possible to observe the impact of these elements. In the high reduction test, the freezing lining is composed of magnesian olivine. Switching to a low reduction process increases the Fe content of the olivine and causes orthopyroxene to appear, making the freeze-lining mechanically and thermally weak. Thermal incidents in the furnace cause significant slag foaming. This foaming is produced by a massive crystallisation of the olivine (due to the drop in temperature), and inducing an exponential increase in viscosity, preventing the free circulation of gases. These foaming episodes are also recorded in the freeze lining and form layers rich in gas bubbles, and therefore mechanically fragile. The slag casted and cooled in a ladle allows the first steps of slag crystallisation to be studied. Olivine is the liquidus phase, with a magnesium-rich core and an iron-enriched rim, contemporaneous with the orthopyroxene phase. The habitus of these phases is very similar to that found in komatiites for the same minerals. FeNi slag is therefore an industrial analogue that is easy to study and compare to komatiites. Keywords: Pyrometallurgy, ferronickel, slag, freeze-lining, characterisation, olivine, mineralogy

Jan 1, 2020

By H. G. Mosenthal

In any mining enterprise, the extraction of the ore from the orebody is of vital importance. This activity is stoping. The gold mines of the Witwatersrand and its extensions in the Klerksdorp, Welkom, and Evander areas form the most extensive goldfield in the world. This address reviews stoping technology and improvements in this technology since the end of the Second World War. Although the reefs that make up the Witwatersrand orebody are uniquely extensive, they are not suited to mechanized mining. This is due to a number of factors such as the depth of the orebody, its narrowness, and the hardness and abrasiveness of the quartzite forming the orebody. The most notable improvements in stoping technology have been brought about because of our beter understanding of rock-engineering problems. This has resulted in beter layouts, stoping geometry, systems of aleviating rock stresses, and support systems. It has enabled the industry, at ever-greater mining depths, to maintain and even increase productivity. These improvements, based on rock engineering, have significantly reduced the accidents due to rockbursts. There are many diferences between today's stoping and that practised forty years ago. Panels are shorter, concentrated mining has superseded bench mining, and the support systems are certainly better. Yet the face scraper is still used extensively, as are pneumatic rock drills. Some innovative stoping systems that have given rise to improved productivity are discussed. Explosives technology has shown spectacular advances during the period under review. Sadly, our local industry has had little to do with this research and development, but has been most adaptive in using the new products on offer.

Jan 1, 1990

By R. A. Govender, A. Mudau

"Rockbursts continue to be a major contributor to mine accidents and cause of damage to mine excavations, particularly at great depth. The problem of rockbursts has also escalated in civil engineering tunnels driven at depth, due to high levels of in situ stress at such depths. Attempts have been made in the past to mitigate the impact of rockbursts, with rock support remaining the ultimate method for providing stability in rockbursting environments. However, records show that, on many occasions, conventional support elements such as rockbolts, wire mesh, shotcrete, and lacing fail to withstand severe rockburst events.Recently, a support method termed ‘sacrificial’ support was proposed as a potential additional method to prevent rockburst damage, based on observations made after rockburst events in a mine. The philosophy behind a sacrificial support system is that, under dynamic loading conditions, support in the form of a liner must fail; leaving behind, undamaged, what was previously supported rock mass. The sacrificial support concept reported herein is applicable in situations where the source (i.e. seismic event) of the rockburst is located remote from where rockburst damage is likely to occur.Spalling tests based on the split Hopkinson pressure bar (SHPB) technique were conducted to study some aspects of dynamic rock fracturing in tension at high strain rates and the role a sacrificial layer plays in combating dynamic rock failure. A single Hopkinson pressure bar, with a long cylindrical intact rock specimen attached at the bar free end, was impacted by a striker on the opposite free end in order to generate a dynamic stress pulse responsible for spall failure upon reflection from the specimen free end. Different liners and liner combinations were then introduced at the specimen free end as support. Such a simple, yet robust, experimental set-up allowed the potential benefits and failure mechanisms associated with sacrificial support under dynamic loading to be demonstrated and compared with ‘sacrificial support’ behaviour observed in real rockburst events in a mine. Analysis of the results revealed that varying liner thickness and mechanical impedance between rock and support liner play a significant role in limiting rockburst damage."

Jan 1, 2016

By J. A. C. Diering

The simulation of mining in faulted ground was recently described by Crouch. His method is extended here to enable a wide range of non-homogeneous mining problems to be modelled. Essentially, the boundary conditions at geological interfaces are incorporated into the overall iterative solution, which is necessary if faults and joints are to be modelled realistically. Two examples are given to demonstrate the method: the mining of coal a seam below a thick dolerite sill, and the mining of a flat reef in close proximity to a vertical fault where hangingwall and footwall rocks have different properties.

Jan 1, 1980

By Z. Tshentu, M. Ngayek, P. Moleko-Boyce

This paper is focused on the development of iridium(IV)-specific quaternary diammonium cations containing electron-donating substituents in the benzyl group of methyl and electron-withdrawing substituents such as triflouromethyl and nitro groups. Quaternary diammonium cations used were tetra- 4-methylbenzyl-1,10-diammonium decane chloride (QuatDMDAMeBnz), tetrabenzyl-1,10- diammoniumdecane chloride (QuatDMDABnz), tetra-3-trifluoromethylbenzyl-1,10- diammoniumdecane chloride (QuatDMDACF3Bnz) and tetra-4-nitrobenzyl-1,10-diammoniumdecane chloride (QuatDMDANO2Bnz). Polyvinylbenzylchloride (PVBC) nanofibres were used as support material, and the functionalised nanofibres were used for separation of Ir(IV) from Rh(III). The sorbent materials were characterised by FTIR and elemental analysis. The efficiency of the quaternary diammonium cations was investigated in a column study under dynamic flow adsorption conditions. Column sorption of [IrCl6]2- and [RhCl5(H2O)]2- on nanofibres functionalised with diammonium cations showed [RhCl5(H2O)]2- was not adsorbed by the sorbent materials while [IrCl6]2- was loaded onto the column. The loading capacities of [IrCl6]2- on the quaternary diammonium sorbent materials increased in the order of F-QuatDMDAMeBnz (3.08 mg/g) < F-QuatDMDABnz (3.39 mg/g) < FQuatDMDACF3Bnz (5.46 mg/g) < F-QuatDMDANO2Bnz (7.10 mg/g). The loading capacity for Ir(IV) with quaternary diammonium cationic nanofibres increased with an increase in the electron-withdrawing nature of the substituent on the quaternising group. The charge delocalising ability of the 4-nitrobenzyl group resulted in the best interaction of the decamethylene diammonium cation with [IrCl6]2-. Keywords: Iridium(IV), rhodium(III), quaternary diammonium, polymer nanofibres, separation

Jan 1, 2020

Most processing plants use automatic samplers to extract ore aliquots along a producing shift. These aliquots are combined into a sample normally sent to lab analysis. The sampler used in the extraction of an aliquot from a continuous flow of ore, i.e. a belt conveyor is required to be both precise and accurate. Surprisingly, not all commercial automatic samplers exhibit these two mandatory properties. Consequently, it is recommended the development of a procedure able to check sample assays reproducibility obtained using a given sampling protocol. This paper uses the Gy’s sampling theory to evaluate coal quality control practices at a washing plant using an automatic sampler extracting coal increments from the ROM feeding belt conveyor. Gy’s theory provides the means to establish the precision limits for a given sampling protocol. Various ore properties are required to be determined in combination with sample mass and mass of the lot sampled. All this information leads to the definition of the precision limit. The precision limit obtained was checked via an experiment mimicking what happens in situations where both miner and consumer use the same and appropriate sampling practices. The experiment demonstrates that sampling theory precision limits were valid if unbiased samplers are used by both parties, i.e. miner and consumer. Keywords: sampling, quality control, coal

Jan 1, 2003

By J. H. Viljoen, C. H. Graan, N. B. Strydom, J. S. Greyson

M J Martinson (Fellow) There is everywhere a lamentable dearth of quantitative information on occupational factors affecting the health and safety of mineworkers, and as the authors point out in their introduction underground lighting is one of the factors which has received little attention in South Africa prior to this welcome paper. Since the authors are to some extent pioneering, it is perhaps inevitable that the paper should raise more questions than it answers, but nevertheless it provides a provisional yardstick against which the need for, and success of, remedial action may be judged. Before commenting in detail on a few aspects of the paper it may be worthwhile to summarize the statutory provisions relating to underground lighting, contained in Chapter 15 of the Regulations promulgated in terms of the Mines and Works Act 27 of 1956. Chapter 15 is entitled `Lighting, Safety Lamps and Contraband&apos;, but in the context of the present paper only four simple regulations need be noted: firstly, no person shall work or travel in an unilluminated part of a mine unless he carries a light ; secondly, adequate stationary lights shall be provided at stations and certain other localities in shafts, winzes and places used for winding; thirdly, mechanically propelled conveyances shall be provided with an effective bright light shining in the direction of travel; and fourthly, moving machinery to which personnel have access shall be illuminated so as to be clearly visible. In all four regulations the standard of lighting to be provided is wholly subjective, and under these circumstances the wide inter- and intra-mine variations in light intensity disclosed by the study are hardly unexpected. Nevertheless the data presented in the authors&apos; Table I raise three points (1) Is there any evidence that the 19 mines visited during the survey are representative of the 45-50 large gold mines in South Africa, or that the shafts selected are representative of the respective mines? Mine managers would scarcely be human were they not to offer better lit working places for the survey; perhaps the use of a scaling factor (e.g. tonnes or transit mineworkers per shift, kilowatts, etc.) for each working place might have helped to put the figures in perspective. (2) Looking at two of the more numerous - and probably important - of the ten working places included in the survey, i.e. stations and tips, one is struck by the marked. skewness of both intensity distributions (with higher frequencies at the lower intensities) and the abrupt cut-off in both cases just short of pitch darkness. Thus, of the 45 meaned readings recorded at 26 tips no less than 19 fall in the lowest range of 0,1-0,9 1m/ft2, with but one tip in total darkness; of 46 readings on 26 stations, 7 fall in the lowest range and none is in total dark-&apos;1 ness. Although a light with an intensity in the range 0,1-0,9 1m/ft2 may satisfy the regulations, it is certainly arguable that in some situations it may be safer to have no light at all. (3) In view of the skewness referred to above, one wonders how meaningful are the `total average light intensities for (the gold mining) industry&apos; quoted in Table I, which the authors compare with illumina¬tion levels found in other countries. Looking at tips again, one notes that 34 out of 45 meaned intensity readings fall below the `total average&apos; for tips, and only 10 exceed it. From this it may be concluded that a few well-lit working places included in a `total average&apos; may conceal a plethora in Stygian gloom. In their discussion the authors make fleeting reference to the effect of underground lighting on productivity and accident rates, without however quoting any South African experience in this regard. Intuitively it seems highly likely that better lighting (in the widest sense) would improve productivity and reduce accidents, and of course it is in these areas that the mining industry might seek economic justification for providing better lighting. It is all the more regrettable therefore that these factors were not studied simultaneously with the present investigation, especially at the Psychological Division of the Human Sciences Laboratory has recently published a study on underground accidents in the gold mining industry without making any reference whatsoever to lighting as a causal factor. In the event neither of the two studies enables us to make an economic appreciation of the situation, and this seems to be a classic example of a situation where some dovetailing of research effort would have greatly increased the useful return. On the effect of lighting on accidents, it may be noted that in 1939 Mickel presented some very interesting statistics to the Mine Managers Association of the Transvaal2, on the improvement in accident rates which followed the random issue of electric cap lamps to native mineworkers in the Hercules Section of E.R.P.M. the previous year. Based on a 12 month trial period in which 250 000 shifts were worked with carbide lamps and (contemporaneously) 307 000 shifts with cap lamps, the incidence of major accidents (at least one day in hospital among men using carbide lamps was 1,15 accidents per 1000 shifts, while the incidence of minor accidents (requiring dressing only) was 11,52 per 1 000 shifts. The corresponding figures for men using cap lamps were 0,79 and 7,31 accidents per 1 000 shifts respectively, representing a reduction of over 30 per cent in both major and minor accidents. While these figures are obviously dated - only men using cap lamps were issued with hard hats, for instance - they do indicate a need for similar research today.

Jan 1, 1971

It is well known that the electrochemical precipitation of metals from aqueous solutions is conducted by the electrochemical standard potential scale. This scale indicates the order of nobility of different metals. Consequently, the standard potential scale shows the nocivity of metals when they are present as impurities in aqueous solutions containing metals. In the hydrometallurgy of cobalt, nickel and zinc are co-deposited during cobalt electrodeposition. Nickel has almost the same chemical, physical and electro-chemical properties as cobalt. Furthermore, the electrochemical potential of nickel is relatively higher than that of cobalt. For this reason nickel is preferentially reduced at the cathode in the presence of cobalt. The presence of small amounts of zinc in solution leads to contamination of cobalt electrowon. In spite of its lower electrochemical potential in comparison to cobalt, zinc is reduced on the cathode because of its higher hydrogen-overvoltage. But, in the hydrometallurgy of zinc, small amounts of cobalt present solution will contaminate the zinc reduced at the cathode. This is easily explained by the fact that cobalt is more electropositive than zinc. The co-deposition mechanism of sulphur at the cathode is not yet well understood. Great attention has to be paid to the mechanism of co-deposition of sulphur during cobalt electrolysis. This work aimed to investigate the influence of some parameters on the co-deposition of nickel, zinc and sulphur during cobalt electrowinning. The following parameters have been investigated: Co2+concentration in the electrolyte, current density and ammonium sulphate (NH4)2SO4additions. It has been established that important nickel, zinc and sulphur co-deposition takes place at relatively low Co2+ concentration in solution. The behaviour of these elements with regard to the cathodic phenomena has been established.

Jan 1, 2005

By M. Abdellatif

The Mintek Thermal Magnesium Process (MTMP) process is based on DC open arc metallothermic smelting of calcined dolomite at normal pressure and temperatures of 1650?1750°C. The process was demonstrated at the pilot scale in the period 2000?2004. Major changes to the MTMP pilot plant were made throughout the development period, which lasted for about five years. The changes included design and installation of new equipment, modification of existing equipment, changes in the metallurgical approach in terms of feed rates, feed recipes, furnace operating temperatures, cold and hot dolime feeding, and steady and systematic changes in the operating procedures and maintenance of the plant. As a result, the performance of the pilot plant gradually improved from one campaign to the next, in terms of availability and in particular the condensation efficiency. In Run 10, the process was successfully demonstrated over an 8-day campaign. Keywords Magnesium, condensation efficiency, DC arc furnace, silicothermic reduction.

Jan 1, 2011

By J. Härkki, R. Hiltunen

The Akanvaara intrusion (surface area 50-55 km2) in Savukoski, Northern Finland, hosts 23 distinct monocumulate layers of small euhedral chromite crystals enclosed in a silicate matrix. These chromitite layers have been divided into four groups; UC (upper chromitite), ULC (uppermost lower chromitite) LC (lower chromitite) and LLC (lowermost lower chromitite). The driving force for investigating the Akanvaara chromites and their reduction behaviour originated from their very low MgO content and relatively low Al2O3 content, giving them a unique quality which based on thermodynamics should facilitate the reduction. Another interesting feature is the high amount of V2O3 present especially in the upper chromitite. The chromites included in this study are from the UC and ULC layers (24.9 and 22.7 per cent Cr2O3 respectively). To obtain a better understanding of the reduction behaviour of Akanvaara samples in comparison to MgO-bearing chromitites, Kemi chromite from the ferrochrome works in Tornio, Finland, was included in the study as a reference sample. The reduction experiments were isothermal in a temperature range of 1000-1400°C in CO atmosphere. Experiments were also conducted with coke addition in pellets and briquettes in CO/Ar/H2 atmospheres. The results of the reduction experiments complemented with microanalytical research indicate that the Akanvaara chromites do reduce to a higher degree than the Kemi chromite. Although the aim was to characterize the mechanisms of reduction focusing on gas-solid reactions, the nature and amount of the gangue minerals was such that a considerable amount of slag phase was formed, which affected the reduction mechanisms extensively. The structural features together with low MgO and high FeO content in both Akanvaara chromites are important factors resulting in better reducibility than in the Kemi chromite.

Jan 1, 2004

By C. J. Fauconnier

The various methods of detecting fires and heatings in collieries are reviewed. The methods described range from those relying principally on human senses to sophisticated methods employing infrared technology to detect traces of the products of oxidation and combustion. The advantages and disadvantages of various gases employed as indicator gases for the early detection of heatings and fires are discussed, and recommendations are made in this regard. The operation of a modern infrared carbon monoxide detection system is described, and the differences in operation and application of telemetric systems, tube-bundle systems, and up cast-shaft monitors are discussed In conclusion, some indication is given of the capital outlay required for typical early-detection systems.

Jan 1, 1981

By M. Yavuz, M. Aksoy, G. Once, M. Iphar

This paper is concerned with a study on the floor heaving in the transport road of the GLI Omerler Colliery in Tuncbilek-Turkey. This gallery has been opened at the depth of 200 m through the formation of marl in 4.6 m width and 3.5 m height in cross section. Floor heaving occurred up to 70 cm in height in maximum and caused severe problems in railway and band-conveyor transportation in the main road. In addition; after the current panel mined out, the shield support units could not be transferred to the new panel because of the deformation of roadway cross-section. The I-profile steel arch supports also sank into the weak floor rock. This situation has been studied by numerical analysis using the FLAC3Dfinite difference program in order to investigate the rock mass behaviour around the opening. The obtained outputs of the numerical model have showed consistency with the amount of floor heaving in the gallery at 200 m depth. A new gallery has been driven at the depth of 115 m and M4 longwall panel has just been taken to production. So as not to encounter with the same problems at 200 m depth, especially longwall face moving problem, various numerical models have been studied in order to keep the floor heaving in the acceptable limits required to move shield support units. Keywords: Floor heave, Numerical modelling, FLAC3D, Failure criteria, RocLab

Jan 1, 2003

By C. A. Young, J. L. Bowden

"Cyclic voltammetry experiments were conducted on copper and chalcopyrite (CuFeS2) in the absence and presence of ethyl xanthate, and compared to mass-balanced EH–pH diagrams. The results for copper duplicate those found in the literature and confirm xanthate chemisorption. However, the, results indicate that chalcopyrite oxidizes to chalcocite (Cu2S) and only afterwards is chemisorption observed due to small currents appearing with xanthate. This phenomenon suggests that the mineral's hydrophobicity is induced by more than dixanthogen and copper xanthate. Hydrophobicity was found to be pH-dependent over a range of alkaline conditions (pH 7–11) at narrow potentials (0 to –200 mV). An EH–pH diagram for chalcopyrite with xanthate is presented to illustrate the conditions under which chalcopyrite would be hydrophobic and thereby more thoroughly explain the results in the literature. IntroductionSulphide minerals are generally concentrated from ores by froth flotation, a physiochemical process that induces separation based on differences in hydrophobicity. In this process, a gas is bubbled into a slurry and the minerals that are hydrophobic (water-repelling) adhere to the bubbles and float to the surface while the hydrophilic (water-loving) minerals remain in the slurry. With sulphide minerals, the hydrophobicity is generally established in two ways. The first is by the addition of a collector, which is a heteropolar molecule with a reactive inorganic head group and an inert organic tail. Typically, the head group bonds with the sulphide mineral surface by adsorbtion either chemically (chemisorption) or physically (physisorption) or by complexing with atoms that solubilize somewhat at the surface (surface precipitation). These interactions leave the organic tail protruding from the mineral surface, thereby inducing hydrophobicity. The second method is by oxidation of the mineral surface to form a layer that is metal-deficient/sulphide-rich or comprised of elemental sulphur. Because the layer mimics naturally hydrophobic sulphur, this enables ‘collectorless’ flotation. Furthermore, this oxidation phenomenon adds an electrochemical dimension to the understanding of sulphide mineral flotation."

Jan 1, 2016

By O. S. Tinkler, K. C. Sole

"Although the first large-scale application of copper recovery by solvent extraction took place in Zambia in the early 1970s, it is only in the last decade that this technology has become widely employed in this part of the world and is now a mainstay unit operation in copper hydrometallurgical flow sheets. The mineralogy of the ores of the African Copperbelt, and hence the characteristics of African leach liquors, differs significantly from those in Chile and the southwestern USA, where copper solvent extraction has had a long and successful history. These differences provide operators, metallurgists, reagent vendors, and engineers with many challenges: new approaches are needed to adapt solvent-extraction technology for successful implementation in this region. This paper examines typical operating practice in the African Copperbelt, discusses differences compared with other parts of the world, and looks at some of the challenges and opportunities presented by these flow sheets. IntroductionFollowing the success of the Rancher’s Bluebird and Bagdad solvent extraction and electrowinning (SX–EW) operations in Arizona in the late 1960s, the Tailings Leach Plant at Chingola, Zambia, became the first large-scale copper SX plant in the world, commissioned in 1974. Despite the equipment design now being outdated, this plant still continues to operate successfully, indicating the versatility and adaptability of this technology. Today, there are some 75 copper SX operations worldwide with cathode production above 10 kt/a. The top ten producers currently account for about 40% of the global 4.3 Mt/a SX–EW copper production. South America (predominantly Chile and Peru) is the largest copper cathodeproducing region, with annual production of some 2 Mt. The Central African Copperbelt (Zambia and the Democratic Republic of Congo (DRC)) is second, with cathode production of 1.2 Mt/a, and North America (USA and Mexico) third, producing close to 0.8 Mt/a Cu (Solvay Cytec data, 2014). The remaining 0.35 Mt/a comes from all other regions combined (referred to as ‘Rest-of-World’)."

Jan 1, 2016

By S. Dalton

The use of coal for power generation is being questioned in many developed countries, as a result of ever-tightening environmental restrictions and concerns about the potential impact of CO2emissions on climate. Coal will continue to be the fuel of choice for new power plants for many developing countries with large percentages of the world?s population, because of its relatively low cost and broad availability. Even in many of the developed countries that are under pressure, coal will continue to fuel large percentages of power due to the installed base. Using coal efficiently and in an environmentally responsible manner will be a challenge for technology and politics in this new century, with the politics played out on the world stage. Coal use will vary in the next few decades, including short-term changes such as the use of more diverse coals and the increased use of efficient ultra-supercritical boilers to longer term ?refining? of coal to produce a range of energy and chemical outputs. This paper outlines EPRI?s Global Coal Initiative aimed at needed developments and demonstrations that support the long-term vision of environmentally sound and economical coal use for power generation, and discusses some of the forces that are driving this development effort.

Jan 1, 2001