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By E. Van Der Lingen

This paper provides a review on coloured gold, platinum, andpalladium intermetallic compounds, and discusses the models thatwere developed to obtain these materials. These compounds have acrystal structure of high symmetry, such as the CaF2or CsClstructures, ensuring distinct electron band structures. Variousexamples of coloured gold, platinum, and palladium intermetalliccompounds are provided. More in-depth discussion is provided on thepurple gold (AuAl2) and yellow platinum-aluminium (PtAl2)compounds with CaF2structure, as well as the purplish-pinkpalladium-indium (PdIn) compound with CsCl structure. Precious metal intermetallic compounds are used in jewellery andprovide a new dimension to design. Some of these compounds havealso found use as barrier coatings on turbine blades for jet engines,and more recently, research has been conducted into their potentialuse as catalysts, electro-catalysts, sensors, capacitors, and fordecorative coatings.

Jan 1, 2014

By M. de Beer

The conversion of carbon-containing feedstock into liquid fuels can proceed via the Fischer-Tropsch synthesis, which is catalysed using supported cobalt catalysts. These catalysts contain noble metals to facilitate the reduction of oxidic cobalt species in the catalyst precursor to metallic cobalt, which is the catalytically active species for the Fischer-Tropsch synthesis. This is thought to occur via hydrogen spillover, i.e. the diffusion of atomic hydrogen over the support from the noble metal to the oxidic cobalt moieties. Noble metals may also affectthe dispersion of cobalt in the active Fischer-Tropsch catalyst. The economic aspect of the addition of noble metals to the Fischer-Tropsch catalyst is discussed.

Jan 1, 2014

By T. Nesane, S. S. Mnyakeni-Moleele, L. C. Murulana

The effectiveness of two synthesized ionic liquids, 1-(benzyloxy)-1-oxopropan-2-aminium 4-methylbenzenesulfonate (1-BOPAMS) and 4-(benzyloxy)-4-oxobutan-1-aminium 4-methylbenzenesulfonate (4-BOBAMS), were evaluated for mild steel corrosion inhibition in 1.0 M hydro-chloric acid solution, using electrochemical impedance spectroscopy (EIS), potentiodynamic polarization (PDP), and gravimetric techniques. Organic moieties responsible for the adsorption process on mild steel surface were investigated using Fourier transform infrared spectroscopy (FTIR). Gravimetric analysis revealed that the inhibition efficiency of 1-BOPAMS and 4-BOBAMS increased with concentration, with maximum inhibition values of 90.32% and 97.91%, respectively, at the highest concentration of the inhibitors. Gibbs free energy (ΔG°ads) values indicated a strong interaction between the mild steel surface and the molecules of the ionic liquids, and that the adsorption process was spontaneous. These values also show that the inhibitive nature of ionic liquids against mild steel corrosion is caused by a mixedtype of adsorption film formed on the steel surface. The Langmuir adsorption isotherm was used to describe the adsorption of ionic liquid molecules onto the mild steel surface. Polarization curves showed that 1-BOPAMS and 4-BOBAMS have a similar effect on both the anodic and cathodic half-reactions, indicating that they prevent the dissolution of mild steel through both physical and chemical process. Nyquist plots were defined by incomplete semicircle capacitive loops, showing that the charge transfer mechanism controls the corrosion of mild steel in acidic solution.

Apr 1, 2021

By R. G. B. Pickering, B. Ebner

The hard rock gold and platinum mines in South Africa have seen little development in the technology of mining. Holes are still drilled in the rock—though the latest water hydraulic rockdrills will complete that task in two minutes compared with the eight hours for hand drilling. Charging up is with far safer explosives and with available initiation systems it is no longer necessary to run up the face lighting each fuse. Movement of rock in the stope is predominantly by means of scraper buckets and scraper winches — these were first introduced in the 1920s.Support is still mainly rock, either in sit pillars or backfill, or alternatively wood. Just as themining process has changed very little it is hardly surprising that the productivity of mining is also relatively unchanged. Major efforts have been made to mechanize the mining process. However, most of the narrow reef mines where ‘trackless mechanized mining’ was introduced in the early 1980s have reverted to conventional mining. In the wider reefs mechanized mining is the option of choice, unfortunately most of the gold and platinum mines have very narrow reefs that are not suitable to’ off the shelf’ equipment. The move to introduce room-and-pillar mining into the narrow UG2reefs is at the expense of mining more waste with the reef, fortunately it appears that appropriate waste and reef separation processes are available. Room-and-pillar mining followed by long holedrilling and blasting of a narrow slot of reef out of the pillars currently appears to offer the best of both worlds—maximum reef recovery with minimum waste dilution. All these mining methods are still dependent on blasting to break the rock. The cyclic nature of mining by blasting places severe constraints on the rate of face advance that can be achieved and consequently the utilization of the invested capital. For narrow reef hard rock mining to break out of these constraints and to really make progress in the 21st century it is necessary to follow the lead of the soft rock mining industry and change the technology of mining. In the underground coal mining industry, coal cutting has been proved to be the most cost-effective solution. In narrow reef hard rock mining the future must be based on the development of non-explosive methods of rock breaking that in turn are integrated into continuous mining systems. Non-explosive so that the mining operation can be conducted on a continuous basis, with nodelays for the removal of blasting fumes. Continuous mining also maximizes the return on the capital invested in developing the mine. Rock cutting will define the stope width and that in turn will be designed to minimize waste dilution; it will also minimize damage to the hanging and create a safer environment. Mechanization makes it possible to automate the mining process and further improve safety by positioning the operators in a safer environment. This paper defines the work to evaluate cutting processes carried out by Sandvik TamrockVoest Alpine. The evaluation was in the UG2 and the pyroxenite and norite typical of the Merensky Platinum Reefs. A novel form of mining machine is proposed that will utilize existing technology. Finally the paper describes the interaction between Sandvik Tamrock Voest Alpineand Lonmin Platinum and the initiation of an agreement to develop and trial a prototype hard rock narrow reef mining machine.

Sep 1, 2001

By H. Wagner

It is a great privilege for me to have been asked to present the closing address to this most successful symposium on backfill in the South African mining industry. When I accepted this invitation I did not fully appreciate the tremendous progress made on most aspects of backfilling in the last few years. The number and quality of papers presented at this symposium bear testimony to the mining industry's determination and success in introducing backfill in underground mines to improve strata control and environmental conditions, and to provide better protection to mineworkers against the hazards of rockbursts and rockfalls. Many contributions demonstrated that these benefits can be achieved at a reasonable cost, and in some instances improved grade and higher rate of face advance result in improved economics. In my closing address I will attempt to summarize the progress made in the various aspects of backfilling underground workings and to identify those areas which, in my opinion, require further attention and effort. I will undertake this review under the following main headings: ? fundamentals of backfilling; ? design of backfill systems; ? operation of backfill systems; ? control of backfill systems; and other aspects of backfill systems.

Jan 1, 1988

By G. Akdogan, M. Tadie, A. Tapfuma

The Southern African region contains abundant gold tailing heaps that are an environmental concern due to dust pollution and chemical contamination of nearby water bodies. Nevertheless, these heaps contain gold with potential to be extracted for financial gain. In this study, the feasibility of extraction of gold from a Witwatersrand gold tailings using glycine was investigated using a two-level full factorial design. Effects of variables such as glycine concentration at levels of 0.5–1.5 M with a solid–liquid ratio of 50–200 g/L were investigated. Although the solid–liquid ratio used in this study was below the industrial norm, it provides a starting point for investigating the applicability of this technology. To enhance the gold extraction, potassium permanganate pretreatment and copper ion addition were adopted. The results revealed that 91.4% gold extraction was achieved after 24 h pre-treatment of the tailings using 2 g/L potassium permanganate followed by subsequent leaching using 1.5 M glycine at pH 12, with 50 g/L solid–liquid ratio and 2 g/L copper ions. Statistically, the solid–liquid ratio and potassium permanganate pre-treatment of the tailings were found to be significant at 95% confidence interval, whilst interactions of copper ions and glycine concentration were significant factors. The work demonstrates that potassium permanganate pre-treatment prior to glycine leaching of low-grade secondary gold resources, such as tailings, can be beneficial. Furthermore, the methodology developed in this work provides an opportunity for further investigation of the recovery of gold hosted in complex mineralogical matrices.

Apr 14, 2024

Contribution by G. Narsimhan It was with much pleasure that I went through D.F. Malan's excellent paper. The following explanatory and critical note has been the result of my deep study of the paper. Although by qualifications and experience I have been a chemical engineer, involved over four decades with problems in the modeling and simulation of transport-limited (heat and mass) chemical reactions in multi-phase systems, I have of and on interested myself in challenging problems in extractive metallurgy. This is the first time I have veered of to continuum mechanics, at the age of 71! Author's reply. I thank Professor Narsimhan for his extensive contribution to the paper. It provided me with the stimulus to derive an extended model. However, there are some inaccuracies in his arguments that need further discussion. Professor Narsimhan's main concern is that there is not a good enough fit between the derived viscoelastic Kelvin model and the experimental data supplied by Güler1 (see Figure l1 on page 218 of my paper). I will therefore explain the choice of a viscoelastic model in more detail and also make specific comments on the arguments given in his contribution.

Jan 1, 1996

By A. G. Journel, CH. Huijbregts

"Centre de Morphologie MathématiqueSpecial problems arise in the evaluation of lateritic-type orebodies (Ni-Co, bauxite) or, more generally, orebodies affected by surface weathering. In this type of orebody, the vertical grade distribution shows, from the hangingwall downwards, a systematic increase followed sometimes by a certain decrease, or vice-versa. These trends appear together with irregular variations from one drill bole to the other.The problem is to characterize a vertical structure that is representative not only of the drill hole but of an entire panel or zone of influence, in order to define mineable thicknesses for given cut-off grades. The vertical grade distribution observed on the drillhole is replaced by a new curve, called 'derive' (drift), which is computed by 'universal kriging'.Once the vertical study is completed, new 'service-variables' are defined in the horizontal plane, such as mineable thickness and the corresponding average grade, economical functions (combination of the grades of several minerals). The local and global estimation of these variables can then be performed by a simple bi-dimensional kriging plan. The global estimation is summed up by several tonnage-cut-off grade curves.The practical valuation according to the above scheme of a New Caledonian lateritic nickel orebody is presented."

Jan 1, 2014

By J. P. Beukes

South Africa holds approximately three-quarters of the world?s viable chromite ore resources and dominates the global production of ferrochrome. Albeit completely unintended, small amounts of Cr(VI) are formed during ferrochrome production. Certain Cr(VI)species are regarded as carcinogenic, hence making the treatment of some ferrochrome waste materials necessary. In this paper, the Cr(VI) treatment strategies currently employed by the South African ferrochrome producers are investigated by means of a literature review and a questionnaire survey. From the discussion, it is evident that various treatment strategies are available to deal with Cr(VI)-containing waste in the ferrochrome industry. However, by far the most commonly applied treatment strategy remains the aqueous reduction of Cr(VI) with ferrous iron. The advantages and the correct application of this strategy, together with the disadvantages and pitfalls, are argued. Innovative improvements on historic practices are also discussed

Jan 1, 2012

By M. Karsten

Platinum mines in the Bushveld Complex are located in areas with a high geothermal gradient. This means that, compared to, for example, gold mines of the Witwatersrand region, high rock temperatures are encountered at relatively shallower depths. Thus air cooling and refrigeration need to be considered at early stages in mine development. Another factor that influences air cooling and ventilation requirements is the use of trackless diesel equipment for loading and hauling broken rock. At shallower depths it will be more cost-effective to locate the cooling and ventilation equipment on surface, whereas for deeper mines the more cost-effective option is to locate cooling equipment underground. For intermediate depths, trade-off studies need to be undertaken to determine the optimum mix between more cooling and less air or more air and less cooling, as well as the optimum location for this cooling equipment. Factors that are taken into account include power requirements, positional efficiency, and additional infrastructure requirements. Traditionally the approach has been to ventilate ?all of the mine all of the time?. However, with increasing power costs and growing pressure to ensure sustainability, this strategy must be modified to apply ventilation and cooling on demand. In other words, both cooling and ventilation equipment should be operated only when needed and should be adjusted to the demand at the time. Obviously there will be areas that will need to be ventilated (and cooled) continuously (for example, for dilution of flammable gases), but this can be built in to the overall strategy. In addition, in devising ventilation on demand (VoD) strategies, occupational criteria relating to exposure to gases, noise, and heat need to be taken into account. This paper discusses options available for implementing energy-efficient ventilation and cooling systems and suggests ways of changing traditional thinking in the overall ventilation and cooling of mines. Efficient ventilation and cooling designs can play a major role in improving sustainability and reducing energy costs and carbon footprint.

Jan 1, 2012

By W. K. Loftus, H. S. Simpson

Discussion I. R. M. Cheston (Visitor): I should like to congratulate the authors on this interesting paper which graphically illustrates the overall effects of the gradual developments in diamond concentrating processes which have taken place over the past few years. These final stages of diamond concentration represent only a minor factor in the cost of diamond production but because of the shortage of highly trained people for this work, any easing of the burden on the sorting staff has an importance far beyond the immediate economic sphere. The search for the solution of problems posed by the economic and social conditions of industry is never-ending. The Diamond Research Laboratory is, even now, carrying out further work to improve still more the operation of general diamond recovery processes as described in the paper. Before looking at some of the latest developments in this field of final recovery, there are a few points arising from the paper on which I would like to comment. On page 321, reference is made to the X-ray sorters originally developed by the DRL. The paper gives a figure of 100 per cent recovery of diamonds from +7 mesh concentrates in two passes through the prototype machine. Not wishing to claim miraculous powers for our group, I would prefer to see this given as virtually 100 per cent recovery of all fluorescing diamonds. Firstly however much care is taken, there is bound to be an occasional operating loss. In the test work, 100 per cent recovery was made on many occasions but this was not always so. Secondly, although most diamonds fluoresce strongly under X-rays, some diamonds only fluoresce weakly. Type IIB diamonds, in fact hardly fluoresce at all, but the incidence of this special type of diamond is very low in most deposits. However, in operating the commercial X-ray machines, there is a certain background level of reflected radiation from other feed particles. Unless the diamond fluorescence is several times greater than this, it is not possible to achieve sufficient sensitivity in ejection. A certain small but variable proportion of diamonds from each deposit is always found to fluoresce too weakly to be recovered by the X-ray machine. Tests have shown, that for the De Beers mines, this proportion is considerably less than I per cent. These diamonds are nearly all dark brown or black in colour and therefore of low value. The degree of fluorescence does not, however, depend entirely on the colour or quality of the diamond and some of the brightest fluorescence comes from the lowest quality of boart diamonds. Investigations into the property of the diamond which causes this low fluorescence are being carried out. On page 322 it is suggested that zircon fluoresces in the same colour spectrum as the diamond. This is not quite accurate. The total light given out by zircon under X-rays is of the same order as that of diamond. However the zircon radiation has a much wider spectrum band than the diamond fluorescence. Reference is also made on page 322 to the effect of selective milling in small laboratory mills on diamonds. Perfect diamonds are very hard and very strong and are extremely difficult to break. Imperfect diamonds, which form the majority of diamonds recovered from most deposits, although hard, can be very brittle. Even under slight impact some of these diamonds may shatter to powder. Milling conditions must therefore be extremely closely controlled to minimize breakage, and even so, some breakage will always occur. As suggested in the paper, the necessary conditions are: the use of small balls, slow speed mills and very limited water addition. Tests elsewhere have suggested that the water content of the pulp in such a mill must be less than 25 per cent by weight of pulp to prevent diamond breakage reaching significant proportions. The skin flotation techniques described on page 323 operate on a very small scale. It is of interest to note that in West Africa a large-scale continuous skin flotation machine is used to recover the fine diamonds. In this operation the feed is dried and, after standing, is mixed with water and fed in a single layer onto a woven phosphor-bronze conveyor belt. This belt runs at a shallow angle into a water bath and as the particles are carried through the air-water-interface, the diamonds float off and over a weir into a collecting box. The bulk of the particles, being wettable, sink to the bottom of the tank and are continuously removed. If treated without prior drying, the diamond recovery is poor. If material is treated immediately after drying, a lot of the gangue particles also float. During standing, in the hot and humid atmosphere of West Africa, it is found that the gangue particles recover their wettability much faster than the diamond particles. Optimum selectivity is obtained after standing for approximately 24 hours. At the DRL we have been experimenting with optical filters to differentiate between the fluorescence of diamonds and zircon. By limiting the light transmission to the fairly narrow range emitted by diamonds, it is

Jan 4, 1970

By M. Steyn

The coordination compound tetrakis(1,3-diphenyl-1,3-propanediona to)zirconium(IV), [Zr(DBM)4], (DBM = dibenzoyl methane/1,3- diphenyl-1,3-propanedionol) was synthesized and characterized by single-crystal X-ray diffraction. [Zr(DBM)4] crystallizes in the monoclinic space group P21/c (a = 24.769(4) Å, b = 10.216(5) Å, c = 19.314(4) Å, and = 101.541(5)°). The coordination geometry around the zirconium atom, formed by the eight O-donating atoms of the four -diketonates, is found to be a near-perfect square antiprism. Crystal packing is stabilized by C?H... interactions throughout the crystal lattice. This structure is a noteworthy example of the application of ambient condition synthesis of zirconium compounds in N,N'-dimethylformamide (DMF) as reaction and crystallization solvent, highlighting the possibilities of aerobic condition metal purification through chelation.

Jan 1, 2014

By van der Merwem J. N.

"Coal pillars are prone to scaling over time, and are progressively reduced in size and consequently subjected to increased load and reduced strength. The effective factor of safety therefore reduces. Databases of failed and stable pillars cannot, therefore, be regarded as definitive.More than a decade has passed since the first attempt to predict the lifespan of pillars. This paper presents a review based on the new database and pillar strength equations. The methodology for the analysis is based on the fact that pillars scale. The ultimate safety factor (defined as the ratio between strength and load at the time of failure) is derived, followed by an estimate of the rate of pillar scaling, which results in an estimate of the expected time of failure.It is suggested that the predicted lifespan should not be used as an absolute indicator of expected pillar life, but rather as an index – the pillar life index (PLI) – to complement the safety factor and related probability of failure when evaluating pillar stability. A pillar with a predicted lifespan of less than 500 years would be regarded as being in danger of imminent collapse, while a lifespan of at least 1 000 years should be required for long-term purposes.The differences between this analysis and that published in 2003 constitute a strong case for regular reviews of all the empirically based stability indicators, namely the safety factor, probability of failure, and the PLI. Reviews at 10 year intervals appear to be required.The constants in the proposed method are valid only for the Witbank No. 1, 2 and 4 seams and the Highveld No. 2 and 4 seams.IntroductionIt is well known that over time, coal pillars are prone to scaling, and thus are progressively reduced in size and consequently subjected to increased load and reduced strength. The effective factor of safety therefore reduces and it is quite possible for failure to occur after a long time interval. In fact, scrutiny of the database of failed pillars indicates that failure occurred after periods ranging from 2 to 49 years, and not immediately after mining.The factor of safety for pillars at the time of failure is not the same as that immediately after mining. The concept of the safety factor is based on the pillar strength obtained by statistical analysis of pillar dimensions immediately, or within a month after mining, when the surveys are conducted. Strictly speaking, the safety factor as commonly used is valid only immediately after mining, and it continues to reduce thereafter. The ideal situation would be to base analyses of pillar strength on pillar dimensions at the time of failure, but that is simply not possible because the real dimensions cannot be measured at that time."

Jan 1, 2016

By Ernst-Ulrich Reuther, Alexander Donmen

Mine ventilation in the German coal mining industry is concerned mainly with problems of thermal environment arising from greater depths and with increasing emission of methane gas. The development of computer techniques over the past ten years for solving these and ancillary problems is reviewed. The emphasis from a planning and control point of view falls mainly on ventilation network calculations, temperature predictions and advance estimates of methane emission.

Jan 1, 2014

By R. Simon, P. Pagé, L. Li, P. Yang

The Mathews-Potvin stability method is widely used in the Canadian mining industry as a starting point to determine the maximum dimensions of mine stopes. However, it cannot be applied to inclined (more frequently encountered) mine stopes without conducting numerical modelling to obtain the stress factor A, defined as a function of the ratio of unconfined compressive strength of intact rock to the induced principal stress on the exposed stope walls. The need to conduct numerical modelling significantly limits the application of the Mathews-Potvin method. In addition, given its empirical nature and main application for preliminary design, it is deemed undesirable to conduct numerical modelling, especially elaborate modelling. Alternatively, theoretical methods can provide a much simpler and quicker way to estimate stresses around stopes and the corresponding stress factors. Over the years, a large number of studies have been conducted to estimate stresses around openings excavated with various crosssections. However, theoretical or graphical solutions remain unavailable for mine stopes that typically consist of horizontal floor and roof, and two parallel inclined walls (hangingwall and footwall). To remedy this situation, a series of numerical simulations is first performed for openings with vertical and inclined walls, including typical stopes commonly encountered in underground mines. A group of empirical solutions is then formulated to estimate the induced principal stresses at the roof centre and mid-height of the stope walls. The validity and predictability of the proposed solution have been verified using additional numerical simulations. The proposed solution can thus be used to calculate stresses and the resultant stress factors A around typical mine stopes with any inclination angle and height to width ratio, under any in-situ stress state, without conducting numerical modelling.

Aug 1, 2021

By H. A. D. Kirsten

The structural competence of mesh- and fibre-reinforced shotcrete was investigated in two series of tests on full-scale panels. It was found that uniformly distributed loading was marginally less demanding than point loading, mesh reinforcement was superior to fibre reinforcement, Dramix fibre and Melt Extract fibre were similar in reinforcing characteristics, and 30 mm to 35 mm was an optimum fibre length. Furthermore, the consistency of the fibre content could not be assured on practical mining sites, and an average fibre content of about 1,5 per cent was the maximum that could be obtained for fibre lengths not exceeding 35 mm. Also, owing to its greater flexibility, shotcrete reinforced with diamond mesh was preferable to fibre-reinforced shotcrete in mining excavations subject to squeezing conditions. In view of the difficulties associated with assuring the consistency of the fibre content in mining applications, it appears that the sophisticated attributes of fibre-reinforced shotcrete can be effectively exploited only in civil-engineering applications in which quality assurance is subject to explicit contractual control. Load-deflection curves, together with measurements of compressive stresses in the surfaces of some of the panels, enabled the ultimate flexural strength of reinforced shotcrete to be rigorously formulated, and rigorous design procedures for shotcrete subject to bending and direct thrust to be developed. The application of the procedures is demonstrated in a practical design case in which the moments and thrusts in a tunnel lining are determined in a numerical analysis of the interaction between the lining, the surrounding rock, and the rock-reinforcing system.

Jan 1, 1992

The two methods for calibrating the parameters K and (alpha) for use in Gy’s equation for the Fundamental Sampling Error (FSE)–, Duplicate Sampling Analysis (DSA) and Segregation Free Analysis (SFA)–, are described in detail. A case study using identical broken reef material from a Witwatersrand-type orebody was calibrated using the DSA and SFA methods and the results compared. Classically, the form of Gy’s equation for the FSE raises the nominal size of fragments given by dN to the power of 3. A later modification of Gy’s formula raises dN to the power (alpha), the latter term being calibrated with the coefficient K in the DSA and SFA methods. The preferred value of for low-grade gold ores used by sampling practitioners in the mining industry is 1.5. A review of calibration experiments for low-grade gold ores using the DSA and SFA methods has produced values of K that vary between 70 and 170 and values of in the range 0.97 to 1.30. The average value for is shown to be 1, rather than 3 as originally proposed in the classic form of Gy’s equation or the industry-preferred 1.5. It is suggested that for low-grade gold-bearing ores the equation for the FSE should raise dN to a power of 1. Such an equation for the variance of the FSE greatly simplifies the characterization of gold ores, now requiring only the calibration of K for a given mass and established fragment size. The implications of the simplified equation for the heterogeneity test are that, provided the fragments have been screened to within a narrow size range, any particular size will return a value for K that is acceptable for use in the sampling nomogram.

Jan 1, 2017

By John Freer

The time 06h30 in midwinter is a little early to set out on an excursion from Johannesburg. Yet this was the hour that members, many accompanied by their wives, started to assemble in the parking lot adjacent to the squash courts at the Wanderers Club, Illovo, for a visit to Premier Diamond Mine at Cullinan. All fifty were checked in and aboard the coach by 07hOO. Coffee and sandwiches, served en route by Doris Gardner and Pam Binstead, were most welcome. IMM awards Given below are details of the trust funds, etc., to which applications are invited for grants, etc., payable in 1985. Application forms, which must be returned to the Secretary before 15th March, 1985, are available on request. Applicants should note that, in general, preference will be given to members of the Institution.

Jan 1, 1984

By C. C. Johnson

In the fourteen years that Rössing Uranium Limited has been in production, the metallurgical operations and controls on the leaching plant have evolved through four basic stages. Initially, the emphasis was placed on the achievement of consistent plant operation by overcoming severe startup difficulties. The second stage involved the attainment of the design operating targets, and also the commissioning of the ferric-leaching reactors in order to achieve a ferric ion concentration of more than 3,0 g/l. Improvements in control then became the priority, with the emphasis on consistently achieving the target concentrations of ferric ions, total iron, and terminal acidity. The latest phase has concentrated on the optimization of costs by means of adjustments to the historically established operating parameters in order to achieve large savings on consumables while maintaining the leaching efficiencies. Apart from the obvious incentive of reducing costs in an inflationary economy and a depressed uranium market, impetus for this work was given by a change in the type of ore from the open pit, which has the effect of reducing the extraction efficiency while increasing the consumption of consumable materials. These problems and their solutions are discussed in detail, and the importance to cost-effective optimization of an accurate, up-to-date cost-reporting structure is stressed.

Jan 1, 1990

By J. Hadjigeorgiou, G. Knox

There is a considerable wealth of information on the behaviour of rockbolts under axial loading conditions both in situ and under controlled laboratory conditions. However, there is comparably less data on the behaviour of rockbolts in shear. Several authors have contributed to the understanding of the behaviour of conventional rockbolts’ performance under shear loading. Much of this data have been gathered through experimental programmes conducted on sections of the rockbolt. Considering the continuous coupling between the rockbolt and borehole interface, the approach is justified. Due to the design premise of energy-absorbing rockbolts with discrete couplings, full scale testing is required to quantify the behaviour of the rockbolts. This paper reviews published results obtained from two testing facilities, which are the SINTEF rockbolt pull test rig in Trondheim, Norway and the Epiroc Combinations Shear and Tensile test rig in Johannesburg, South Africa, both capable of conducting full scale testing of rockbolts. Based on this high-quality database it was possible to identify three distinct responses of rockbolts in shear. The behaviour under load is controlled by the presence or absence of an encapsulation medium (resin/cement grout), the reinforcement mechanism, and whether the rockbolt is continuously or discretely captured. As a reference the results are compared to the performance in shear of conventional grouted rockbolts.

Nov 24, 2025