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By B. K. Loveday

There are certain ores that can be treated without upgrading (e.g., gold and uranium ores), and upgrading needs to be considered only for fairly low-grade ores of this type. If the grade is very low, treatment will not be economic even with upgrading, and there is therefore a limited range of ore grades for which upgrading is suitable.

Jan 1, 1980

By P. R. Labrum, H. A. D. Kirsten

The literature on shotcrete is reviewed with the object of identifying aspects that have not been addressed specifically in terms of its structural performance and manufacture as applied to underground mines. The programme and results of a series of full-scale tests on mesh- and fibre-reinforced shotcrete panels are presented, from which it is concluded that fibre-reinforced shotcrete is competent, especialy at large deformations, and is equivalent in structural performance to mesh-reinforced shotcrete for a fibre content of 3 per cent by mass: furthermore, that the support design systems developed for mesh-reinforced shotcrete are directly applicable to fibre-reinforced shotcrete. The Q-NATM system for the design of temporary tunnel support is presented briefly as one such system. It provides for the integrity of the excavation to be considered in terms of its competence and displacement determinateness, and represents the first systematic procedure for the design of temporary tunnel support. Finally, the support designed on the basis of the Q-NATM system for the main and cross-cut tunnels on the production level in the proposed BA5 block cave at Premier Mine is presented as a practical example of the design of fibre-reinforced shotcrete as part of an overall support system.

Jan 1, 1990

By S. L. Crouch

This paper describes a numerical procedure for the computation of the displacements and stresses induced by the mining of a tabular ore deposit in faulted ground. The numerical procedure is based on the displacement discontinuity method, which can be regarded as an application of Salamon's 'face element principle' to cases in which the reef plane is faulted into several sections, or even folded in an arbitrary way. An element of a fault is modelled in the same manner as an element of the reef, except that a fault element is allowed to undergo permanent inelastic deformation (i.e., slip) if the stresses transmitted through the fault-filling material satisfy a Mohr-Coulomb yield condition. An iterative technique for the incorporation of a Mohr-Coulomb yield condition in the general displacement discontinuity method is explained. This technique is illustrated by means of a simple example involving the mining of a flat-lying reef intersected by a fault. The displacements and stresses at selected locations along the fault are examined as the reef is mined, element by element. A consideration of the energy changes reveals that the occurrence of slip along a fault leads to a sudden increase in the energy release rate due to mining as the excavation intersects the fault. The energy release rate then falls off abruptly to a value less than that which would occur in the absence of any slip, and gradually builds up to re-establish the conditions that would exist in unfaulted ground.

Jan 1, 1979

By J-C. Lv, S. Zhao, F-L. Zhang, T. Yan

"To eliminate nozzle clogging during continuous casting of ship plate steel and reduce the calcium addition required, the composition of CaO-MgOAl2O3 inclusions in the liquid phase region must be controlled. Thermodynamic calculations using FactSage show that when the balanced Al content is 0.03 mass %, aCa in the liquid steel is 1–10 ppm, and aMg is 0.01–0.3 ppm; and when the balanced Al content is 0.06 mass %, aCa in the liquid steel is 1–30 ppm, and aMg is 0.05–0.9 ppm. The CaO-Al2O3-CaS inclusions can be controlled in the liquid region below 1600°C at aCa values of 5 ppm, 15 ppm, and 30 ppm, and the balanced Al content should be controlled in the range of 0.001–0.050, 0.005–0.250, and 0.010–0.50 mass %, respectively. Based on theoretical analysis, the refining slag composition should be controlled to 45–53 mass % CaO, 30–37 mass % Al2O3, 6–12 mass % SiO2, and 6–8 mass % MgO. Optimization of the slag system and steel composition can reduce, or even eliminate, the requirement for calcium treatment. The total oxygen content in the slab can be controlled to 15–21 ppm, with typical micro-inclusions ??10 ??m in size, with compositions in the CaO-Al2O3-MgO and CaO-Al2O3-CaS systems. Introduction With the rapid growth in steel output in China and the changes in the macroeconomic environment, competition in the steel industry has become more intense. The technology for economical cleaning of high-grade steel is important for improving product quality and reducing production costs. However, the functions of refining equipment have not been fully taken into consideration, and the combination of refining methods has not been optimized completely. These problems have become a bottleneck, restricting modern steel enterprises from developing an economic clean steel production process. Ship plate steel produced by the converter long process has been selected as the example in this study. The composition of the steel is shown in Table I. This steel is a high-strength low-alloy steel for rolling media and heavy plates, mainly used for ocean-going vessels and other steel structures. The carbon content in the steel is far from the peritectic point, and the crack sensitivity of the slab surface is weak, so impurities and gas content are strictly limited. Therefore, the process route BOF (basic oxygen furnace) – LF (ladle furnace) – CC (continuous casting) needs to be studied."

Jan 1, 2018

By W. Bleloch

Published in the Journal, September 1970, and presented at a colloquium on 'Future of the Steel Industry', 17th March, 1971. Errata (a) The final sentence in the penultimate paragraph, left hand column p. 34 of the text should read: 'In the hot regions of the bosh and stack the C+CO2=2CO reaction proceeds to CO from coke carbon and carbon dioxide generated by reduction of iron oxides by CO (the so-called solution loss of carbon)'. The italicised words have been omitted in the text. (b) Temperature in Fig. 3 is K not C. (c) In Fig. 9, connection from top plate of column 10 to heat exchanger 9 has been omitted. DISCUSSION J. B. BEEDON AND C. J. JONKER* Dr Bleloch's paper has outlined the advantages of recycling stack-gas of low nitrogen content in the blast furnace. In the first instance it would appear to be extremely beneficial in controlling the hearth temperature and allowing a 100 per cent oxygen-blast to be used. On a second point however we have to disagree. It is claimed that heat is liberated in the combustion zones by the oxidation of CO to CO2 near the tuyeres and in the subsequent reduction of the CO2 by CO. The author has hypothesized as follows, and we quote: 'The C+CO2=2 CO reaction in the outer combustion zone at about 1 700°C is no longer endothermic due to loss of stability of CO2 and its consequent dissociation into CO and free oxygen.' This statement requires closer examination. If we consider the individual events occurring with the CO in the recycled topgas, with the oxygen in the blast and with the coke in the burden, as done by Dr Bleloch, the following heat considerations emerge. All heats of reaction are calculated from the enthalpy values of each reactant and product, as well as the temperature dependence of their molar heat capacities. (a) At the tuyeres: CO+1/2 O2-+CO2 . . . . . (1) CO, CO2 at 500°C; O2 at 25°C. ?H reaction = -65 000 cals/mole CO The reaction (1) is exothermic. (b) In the outer combustion zone: CO2+C?2 CO . . . . . . . . (2) C, CO at 1 700°C; CO2 at 500 0C. ?H reaction = -56 000 cals/mole CO Reaction (2) is endothermic. The overall heat gain of reactions (1) and (2) is ?H=+56 000 - 65 000= -9000 cals/mole CO i.e. the overall reaction is slightly exothermic, the amount of heat liberated is however smaller than suggested by Dr Bleloch. (c) The dissociation CO2+CO+1/2 O2 . . . . . (3) is a highly endothermic reaction at 1 700 0C, requiring 82 000 cals/mole CO2 to proceed to completion. Furthermore, this reaction (3) is energetically not favoured to proceed, as can be observed from the free energy-temperature relationship of figure (3), which favours the formation of CO2 at 1 700 0C. Hence direct dissociation of CO2 at 1 700 0C appears to be unlikely.* The concept of recycling topgas is however still intriguing from many other points of view. Apart from providing a very efficient method of hearth temperature control, it will create a much higher CO partial pressure in the stack. Whether the rate of indirect reduction in the stack will be proportional to this partial pressure or not is un-

Jan 9, 1970

Solvent loading: It is agreed that it is advantageous, as far as reagent cost is concerned, to operate at maximum solvent loading. However, owing to the fact that no reliable automatic solvent flow controller is yet available, control tends to be erratic with the result that there is risk of losing uranium. As can be seen from Mr Tunley's plot, the graph representing total ammonia cost is fairly flat between loadings of 3.5 and 4.0 gpl, while the graph representing uranium loss in the raffinate is steep. It is, therefore, safer to operate slightly below optimum loading in the absence of absolute control. Should the purity of the final product gain importance, operation at optimum loading or even slightly above, may be advantageous. Solvent losses and clarification of pregnant solution: The cost of replacing solvent, as indicated by Mr Carman, represents a high percentage of the overall Purlex costs. It is, therefore, agreed that, amongst other factors, efficient clarification of the pregnant solution and mixer-settler design have an important bearing on the cost. However, as at Blyvooruitzicht, the Buffelsfontein plant can tolerate a suspended solids content of 50 ppm without serious deterioration in efficiency. An increase in the suspended solids content from 15 to 50 ppm normally accounts for an increase of approximately 1 ppm in amine loss. Clarification of acidic pregnant solutions is much more difficult than that of alkaline solutions. Sand clarifiers are being used at Buffelsfontein with moderate success. Clarified solutions containing less than 20 ppm of solids in suspension can be produced, depending on the quantity and type of flocculant used in the filtration process, the thickness and the grading analysis of the sand beds and the quality of supervision. An on-stream duty of approximately 0.15 gallon per sq ft of surface area is obtained. Hydraulic handling of clarifier sand, while cleaning, may effect the efficiency of a clarifier, as size segregation of sand particles may lead to inefficient "pockets" in a sand bed. Control of the solution level in the sand clarifier also affects the quality of the effluent. Exposure of the sand bed invariably causes a cloudy effluent. Recycling of the final and initial solution discharged before and after skimming the clarifier sand improves the clarification efficiency significantly. Post-precipitation of silica and possibly sulphate salts, after clarification, may give a false impression of inefficient clarification. This phenomenon is more pronounced when treating hot supersaturated solutions. Ageing and cooling of the solution before clarification assist in stabilising the solution and reduce precipitation after clarification. The maintenance cost of sand clarifiers at Buffelsfontein is in the order of 0.35 cent per metric ton of solution treated. A redundant ion-exchange column has been tested successfully, either as a primary or a secondary clarifier, depending on the quality of the filtrate treated. The conventional pebble and sand beds in the column have been retained. Operation was enhanced by the addition of a two-foot deep resin bed as a filtering medium. The suspended solids content was reduced to less than 5 ppm. Mixer-settlers: Mixer-settler design is a controversial subject but, in view of the high costs of solvent compared to relatively low pumping costs, it is believed that mixing and pumping functions should be separated. Separate facilities for pumping and mixing not only give greater operational flexibility but also permit eventual changes in the flow circuit as technology develops. Scrub technique: Unlike the pilot plant, where the crud was carried away in the aqueous phase, crud is carried over with the organic phase in the full-scale plant. This carry-over of crud into the strip bank aggravates crud formation in the first strip stage. This phenomenon may result from a slight difference in settler design and/or a variation in the relative phase depths in the two plants. Bypassing of the scrub aqueous solution from the fourth to the second stage was introduced in order to avoid dilution and partial neutralisation of the acid added to the third stage, as it is known that a low pH value enhances the transfer of iron into the aqueous phase. The low aqueous flowrate in the third scrubbing stage was overcome by introducing an aqueous phase recycle. Gleaning: Mr MacDonald's contribution on "gleaning" was most informative. This process for recovering solvent from raffinate seems to show at least as much promise as coalescers and other methods employed to date. Similar tests conducted at Buffelsfontein pilot plant gave 60 per cent amine recovery from raffinate containing 10 ppm amine. However, the circuit was not operated at equilibrium and recovery may be expected to drop as the amine concentration of the "Gleaner" circuit increases.

By P. Smit

Health hazards of South African . mIne water by P. SMIT, M.B., Ch.B., D.T.M.H., D.P.H. (Visitor) The water supply in South African mines is not as good as most mining engineers would like to believe, nor is sufficient care taken to ensure the safe dispensing, transportation, and consumption of water. As long as cases of dysentery and typhoid continue to occur in large numbers, it can be assumed only that the water supplies are not safe, or that the water-dispensing methods are poor. A dual water supply is a real health hazard. The ideal water supply is a single safe source for human consumption and industrial use.

Jan 4, 1977

By A. L. Hinde, P. J. D. Lloyd, D. Hebden

The philosophy behind the design of a rougher cell for the handling of coarse feed and for the maximization of recovery within a given time is presented. The application of this philosophy in cells of 22-litre, 320-litre, and 3 m3 capacity is outlined, and the results of the process testing of these cells are given. In this testing, particular accent was placed on the recovery of water as a guide to the recovery of the finest size of particles, and the rate of recovery of various mineral species. It is concluded that a promising concept has been identified, but that further testing is needed for the firmer establishment of optimal operational parameters in a production environment.

Jan 1, 1983

By C. J. Alberts

The perception exists in many organizations that, once an energy management system is employed, there need not be any further adjustments to such a system. This is not true, as the field of energy management is a dynamic field, and constant feedback and adjustments are needed to enhance the efficiency of a system to the point that substantial pay-back can be obtained. It is stressed that for any energy management system to be successful, careful planning should permeate through the organization, from the top level decision-makers to the workplace. This paper describes a dynamic technique by which energy management strategies can be evaluated by viewing the effects of the strategies on the organization as a whole.

Jan 1, 1997

Written contribution by J.F. Curtist The paper is an interesting exercise in the application of mathematics to a theoretical situation presumed to oc¬cur underground. However, some of the assumptions on which the calculations are based are unlikely to occur in any actual mining situation. Assumptions made in the paper Objections to these assumptions are noted seriatim. (1) `it is assumed that any loading-rate effect will be small and can be ignored' (p. 3). A loading rate of the order of 50 kN-12 per cent of the maximum load-is customary to prevent the props being dislodged by the blast. (2) `it is unlikely that all the props will be installed with their full stroke available, but for the purpose of this analysis, it is assumed that this is the case' (p. 3). A more reasonable assumption would be that, with the most careful selection of prop length and the most rigid control of stope width, the available stroke length would on average be reduced by the order of 70 mm on installation-20 per cent of full stroke. (3) `each support unit supports a block of 3,3 m2 bound¬ed by vertical fractures with zero cohesion that are parallel and at right-angles to the stope face respec¬tively' (p. 5). Joughin and Jager have shown this to be an unwar¬ranted assumption in respect of the positioning, the direction, and the dip of the fractures present in a stope in rockburst conditions. If, as assumed, there was `zero cohesion' between such blocks, the removal of a hydraulic prop in row 3 supporting such a block would ensure that it fell out before a profile prop could be installed in the same position. (4) `the height of each of these blocks is 3,0 m, the volume of each block is 9,9M3, the mass of each block is 27 225 kg' (p. 5). The concept of individual props being affected by dis¬crete blocks of known and uniform dimensions of relatively low ground velocities (less than or equal to 2,16 m/s) acting in line with the axis of the prop is one that lends itself to neat mathematical calcula¬tions. It does not, however, bear much similarity to the actual conditions that are exposed underground in excavations where rockbursts occur. Ortlepp and co-workers concluded that `the impulse of the shockwave can cause the displacement of the

Jan 1, 1991

On 8th August, 1973, 57 members of the Institute visited Prieska Copper Mine at Copperton in the North Western Cape. After tea, the group was welcomed by the Management of the Mine, and Mr Grobler, the General Manager, gave an introductory talk. After they had been taken on a bus tour of the property, the members split up into groups and proceeded underground or to the concentrator. Of special interest to the mining members were the rock-handling methods in use at the mine. The stoping method that has been adopted is the continuous, long-hole sub-level open stoping system. The broken ore is moved with rubbertyred front-end loaders from the stope drawpoints to the ore passes. On the main tramming levels, the ore is collected by electric locomotives and discharged into the main shaft system at the bottom. At the shaft ore-pass system, a gyratory crusher that can accept rocks up to 1 m in diameter has been installed. The crushed product (minus 150 mm in diameter) will then be hoisted by a Koepe winder, which has a rated capacity of 900 tons per hour. The mine is also served by a declined roadway from surface, which facilitates the movement of heavy equipment. The most striking aspect is the productivity that can be achieved by the use of a trackless-mining system. For the production of 230 000 tons per month, it is estimated that 6 stopers and 7 developers will be required. The total labour force will be approximately 2500-2000 Bantu, 200 Coloured, and 300 Europeans. Training of maintenance and operating personnel is of the utmost importance, and it was interesting to note the extent to which non-White labour is being used in these occupations. Lunch was served at the single quarters, after which the group proceeded to the airfield for the return trip to Johannesburg. On behalf of the visitors, Professor de V. Lambrechts thanked Mr Grobler and his staff for the most interesting morning and for their hospitality,

Jan 10, 1973

By R. G. B. Pickering

Conventional mining is an excellent match of mining method and equipment. The result is guaranteed performance. Change in the form of different technology requires change in mining method until the best match is achieved. This paper covers the development of technology and mining method that has occurred in the last five years to arrive at the objective of successful mechanized mining in narrow reef platinum deposits of the Bushveld Complex.

Jan 1, 2004

By J. H. Lazarus

The evolution of backfill-distribution systems for deep gold mines is examined. The status of current technology for the design of backfill-distribution systems is considered with reference to pipeline pressure gradients, pipeline wear rates, and slurry flow regimes. The effect of an increase in the relative density of the slurry on the performance of backfill-distribution systems is illustrated. Future trends regarding the design of backfill-distribution systems are discussed.

Jan 1, 1993

By O. Korhonen, T. Leskinen

The use of large OK-16 flotation machines at the concentrators of Outokumpu Oy by o. KORHONEN and T. LESKINEN(Visitors) The development ofthe OK-16 flotation machine is described briefly, and an account is given of its use at various concentration plants in Finland. The benefits arising from the use of larger flotation machines are listed, and the technical specification of the OK-16 machine is detailed in an appendix.

Jan 6, 1976

By C. Ringas, F. P. A. Robinson

Various theories proposed to account for microbial corrosion are reviewed as a background to a better understanding of this type of corrosion in iron-based alloys. Experimental evidence is presented to show that even stainless-steel alloys are prone to microbial attack by sulphate-reducing bacteria. The corrosive attack by the bacteria is due primarily to the hydrogen sulphide that is produced as part of their metabolic processes.

Jan 1, 1987

By А. V. Aksenov, S. G. Rybkin, А. Y. Senchenko

Recovered activated carbon fines result from processing of ore and concentrates via cyanide leaching and sorption of valuable metals on carbon at gold-mining plants. This intermediate product is characterised by the particles sized smaller than 0.6 mm and gold grade up to 300-500 g/t. Processing of carbon fines with recovery of valuable metals is not available at many plants due to the lack of optimal technology. The paper describes the testwork and development of an efficient technology to process carbon fines with the recovery of valuable metals into the target gold alloy. This innovative technology involves pelletising of fine activated carbon particles using an organic binder, drying and further firing of pellets in a grate fired furnace at 800-900 °С. The produced ash is subject to sintering with sodium nitrate and hydroxide at 600-700°С. The sintered ash is then leached in water. The insoluble residue containing gold and silver is removed by filtration, dried and smelted to produce gold alloy. Water solution containing sulphur and arsenic is subject to detoxification. Processing of recovered carbon fines according to the developed technology implies the use of standard full-scale equipment. The offered method to treat gold-bearing carbon fines was patented by the Eurasian Patent Office. Key words: Fine activated carbon particles, valuable metals, pelletizing, firing, ash, sintering, leaching, smelting

Jan 1, 2020

By L. A. Vermeulen, D. D. Howat

New distribution functions are derived to describe the size distribution of grinding elements in ball mils. The formulations are based on the assumption that abrasive as well as impactive interactions occur during bal milling-an assumption that is supported by a large body of experimental evidence. It is shown that the functions can be used in the estimation of the magnitudes of abrasive and impactive components in the total rate of bal wear. The relative magnitudes of the wear components provide a basis for optimization of the chemical and metallurgical properties of the balls in a given milling situation. However, it is contended that these quantities are also useful indicators of the relative intensities of the abrasive and impactive interactions that are operative in the size reduction of mineral particles within ball mils. These quantities can be determined for any industrial ball mil, and their magnitudes provide practical guidelines for mil operation. The theory is used in an analysis of samples of ball charges from two industrial ball mils, and it is also applied to al the data on bal-size distributions found in the literature. The qualitative correlation between the calculated values of the wear components and the reported operating conditions is good for a variety of industrial ball mils, in conformity with the hypothesis that the relative magnitudes of the wear components are related to the milling conditions.

Jan 1, 1986

Project evaluation is completed through individual studies covering technical and financial aspects of the mine to market. These studies are completed at various levels of detail depending on where the project evaluation is in the project's life cycle. The competency of the studies varies in relation to the quality of data, level of the skills employed, control, management and reporting structure. The focus of this paper is managing elements of the study, looking at the cost of opportunity to influence the project, against the value added. Selection of the study manager and study teams is central to getting an integrated project evaluation addressing the correct issues completed on time. Retention of key personnel can reduce rework which, in turn, can reduce time and cost impacts. It is important to have a work breakdown structure at the outset so that the study teams have a common numbering and naming system that will integrate areas, equipment, budgets estimates and accounts. Having set up the evaluation groups and provided training in the study procedures, understanding the scope and schedule are vital elements in producing a project evaluation that addresses the opportunities and the risks of the final project. Arriving at an accurate capital budget requires integration of the research controlled through common systems. Project evaluation process must also address how the project is to be delivered.

Jan 1, 2007

By J. van der Walt, A. Whillier

Reference is made to the approach to mine cooling in which the service water is used as one of the means of distributing refrigeration, and attention is drawn to various engineering disciplines or functions necessary for effective distribution and reticulation of cold water from refrigeration plant. The demand for service water is intermittent, the peak flow-rates being typically two to three times the average flowrate. The refrigeration plant, on the other hand, must preferably be operated steadily throughout the day and night, with the result that a large storage dam for cold water must be incorporated into each system in order to even out the surge in demand for water. These dams should be operated as constant-temperature, variable-volume reservoirs. The conventional cascade system involving dams on each level is not the best for sending water down mines, and schemes are described that avoid the need for such dams. Reference is made to various tests that were carried out, and to several large systems that were installed. Management and organizational problems that could arise as a result of new technological developments in mine cooling are referred to, and cost figures are quoted to indicate the large potential savings, In addition to the improved environmental conditions, that can be achieved through the use of integrated water systems.

Jan 1, 1979

By D. D. Hawat

The locations of potential reserves of chromite ores throughout the world are discussed. Reference is made to the marked changes in international trade in chromite ores during the past ten to fifteen years, and to the change in locality of the production of ferochromium aloys, which is curently largely carried out in the country of origin of the ores. The influence of these changes has been very profound in South Africa, which has the largest smelting capacity and the largest production of ferochromium aloys in the world. A proportion of South African ore is stil exported to meet the demands for the production of chromium chemicals and refractories, including a large tonnage of foundry sands. Ferochromium smelters in South Africa have a capacity of virtualy 1 Mt of aloy per annum. Nearly al this development has taken place since the introduction of the argon-oxygen decarburization (A.O.D.) process for the smelting of stainless steels. Formerly, expensive Iow-carbon ferochromium aloys were required as raw materials, but the AO.D. process, which is being used world wide, permits the use of high-carbon ferrochromium, the main product of the smelters in this country. The production in South Africa of chromium-containing corrosion- and heat-resistant steels is stil on a relatively smal scale, but serious atention is being paid to the exploitation of low-carbon weldable feritic chromium steel.

Jan 1, 1986