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By Jürgen Antrekowitsch, Harald Osthof, Stefan Steinlechner

"In the case of feasibility assessments or feasibility studies, an important point of discussion is often how to calculate the capital (CAPEX) and operating expenditure (OPEX) of a plant. Generally, there is also the request to know how the CAPEX varies when the size of the plant is increased. Nowadays, many databases give an indication of specific equipment costs in relation to the production rate. These data are based on cost information from equipment suppliers and have to be continuously updated by the owner of the database. The approach presented here is a model which allows sizing of process equipment, for example a leach tank, based on process data in combination with the calculation of mechanical properties. Furthermore, it calculates the CAPEX based on construction material cost and adequate machining surcharges. Thus, a single cost model for a leach tank is adequate for both copper and other hydrometallurgical plants and stays valid with changing construction material prices. To summarize, the advantage of this model is the more versatile approach and thus the flexibility due to the inclusion of design aspects, process and mechanical engineering aspects and an updatable database. The database, which is in regard to material cost and is the basis for the cost estimation, is easily updatable from commonly accessible data. This model thus approaches the procedures of an engineering company when designing a plant and executing a cost calculation. INTRODUCTION Hydrometallurgy nowadays plays an important role in processing of primary and secondary resources. In copper metallurgy, hydrometallurgy accounts for roughly 20 % of worldwide primary production, especially in the case of oxide or low copper-containing ores. Zinc smelters changed their method of production from a pyrometallurgical-dominated concept decades ago to mainly hydrometallurgical routes today. Gold production processes using leaching, especially, have a long history without any development which would not be competitive if not based on hydrometallurgical concepts (Pawlek, 1983; Winnacker-Küchler, 2006)."

Jan 1, 2015

By L. von Ketelhodt

During the period October 2003 to June 2004, a containerized optical sorting pilot plant was operated at Kloof Gold Mine on a waste rock dump. In this paper, we describe the test work and its results leading to the decision to operate this pilot plant. This paper presents the operational data and financial figures achieved during this project, in particular, for the month of June 2004 when the plant was run at full capacity on three shifts, 24 hour operation. These results are also extrapolated and presented to show the profitability of such an operation at today?s gold price levels and cost structure.

Jan 1, 2009

By Y. Qin, L. Han, P. Gao, Y. Han, Y. Lia

In the present study, an innovative comminution technology of high-voltage pulse discharge was proposed for galena to weaken the mechanical property. The results of pre-treatment experiments indicated that the optimal operating conditions of spherical gap spacing 25 mm, pulse numbers 120, and loading voltage 25 kV were obtained, where the relative grindability of pretreated products was improved significantly, although with the prolonging of grinding time, the superiority of high-voltage electric pulses was gradually weakened. The galena ores were subjected to a high-voltage pulse discharge, and then the crack generating along the boundary between galena and gangue was observed under an optical microscope, and the surface burning which appeared in the ores. The scanning electron microscope (SEM) results indicated that the pretreated products were in a state of melting and then condensation, accompanied by some pore formation. Pulsed discharge pre-treatment not only produced more micro-cracks and cracks in the sample but also extended the original crack. Keywords: High-voltage pulse discharge; galena; relative grindability; micro-crack

Jan 1, 2020

By A. C. Mulholland

The electrical control of three-electrode submerged-arc furnaces suffers from a number of complexities, mostly due to the nature of the furnace electrical circuit providing power to the furnace where the three electrodes are interconnected within the furnace through the molten metal bath. This gives rise to what is known as the interaction effect 2, where variations in one electrode?s current can cause comparable changes to the currents in the other electrodes, particularly in larger furnaces with low power factor. Resistance-based control of the electrode penetration has largely alleviated these problems, since the resistance encountered by each electrode is predominantly dictated by the length and conductivity of the current path from the electrode?s tip to the molten metal bath, which acts as the three-phase circuit?s floating neutral point. Hence resistance changes due to tip position or conductivity changes beneath one electrode do not affect the resistances beneath the others, effectively decoupling the control of the individual electrodes1. Although resistance-based control is therefore generally accepted as superior to current-based control for the regulation of submerged-arc furnace electrode penetration, a few furnace operators still prefer to use current control under specific furnace conditions. This paper presents the results of analysing the performance of both current and resistance-based control in typical scenarios encountered on industrial furnaces, taking into account a number of factors including electrode penetration, power distribution, efficiency and asymmetry of the electrode currents. In order to accomplish this, three typically encountered scenarios were simulated. The results obtained show that resistance control provides more benefit in all cases, however, the uneven electrode current distribution generated by resistance control when electrodes are on top stops may cause some concern if baking of the electrodes is required in this scenario.

Jan 1, 2009

By E. Lwamba

The aim of this study was to examine the effect of grain size distribution control of the raw material mixture on the permeability of the green sinter bed and the properties of the produced sinter. This was achieved by evaluating the granulation characteristics of the sinter mixture (moisture content, granulation time and mean granule diameter) in terms of its green bed permeability, and evaluating the productivity of the sinter bed, the coke rate, tumble index (TI), abrasion index (AI), reduction disintegration index(RDI) and reducibility (RI) of the produced sinter. The raw sinter mixtures contained combinations of Thabazimbi iron ore, Sishen iron ore, coke, lime and return fines. The grain size distributions were controlled by removing the ?0.5 mm size fractions of the return fines and coke, and the ?1 mm size fraction of lime. Of the examined mixtures, a mixture of 20 mass% Thabazimbi iron ore, 80 mass% Sishen iron ore and fluxes where the coke, lime and return fines were all sized had the highest granulation effectiveness and permeability. The sintering properties of the mixtures in which the grain size distributions were controlled, were very similar for all the mixtures, but superior to the base case mixture in which the grain size distribution was not controlled. Keywords: Thabazimbi iron ore; Sishen iron ore; granulation, permeability, grain size distribution; sinter.

Jan 1, 2008

By L. H. Lindstad

The paper reviews the tapping process and the major wear mechanisms on the tap-hole, including wear from tapping tools. Measured temperatures in different tap-holes are related to choice of binder and material for repair pastes. The strength and disadvantage of carbon and silicon carbide materials are compared. Normal daily and weekly maintenance of tapholes is described, together with the major repairs that are done during longer furnace shutdowns. The objective is to show that the type of repair and repair material needs to be adapted to each individual furnace.

Jan 1, 2014

By P H. Ferreira

The mechanisation of mine development is becoming increasingly important in our South African and Southern African mines with increased emphasis on safe and efficient mining. Raise, blind and shaft boring offer mechanised solutions for excavations such as:- [reef raises stope boxholes and shaft ore passes ventilation passes and shafts chairlifts large vertical hoisting shafts (in excess of 1200 metres in depth) angled hoisting shafts (from small to large diameter bored holes)] These techniques will be discussed with reference to relevant applications and the author?s comparison between conventional drill and blast techniques.

Jan 1, 2005

By D. F. Malan

Considerable practical knowledge and effective use of empirical evidence is required in the design of tunnel support systems. In many instances, it is unclear how individual support elements interact with fractured rock and which elements of the support system are critical in controlling and containing the movement of material proximate to the excavation. The present paper outlines a displacement discontinuity approach that allows the coupling of surface constraints, in the form of liner or mesh material, to an explicit representation of time-dependent rock discontinuity creep movements. The method is illustrated for a simple square-shaped tunnel profile in which the fracture zone is mobilized by an imposed stress field. A crucial aspect of the model is the computational treatment of the coupling between the liner material and there presentation of the fractured rock mass. This allows the response of the liner material to be assessed quantitatively in terms of the possible reduction of movement in the rock mass. The importance of installing the liner timeously after the excavation is formed is illustrated in a specific example. A number of computational difficulties in simulating liner-rock interaction mechanisms are identified.

Jan 1, 2008

By Y. Cao, C. Li, A. Song

Bubble size plays a key role in determining flotation efficiency. Bubbles can be formed by passing air through pores. Bubble detachment from a pore is driven by lifting forces including buoyancy force and momentum force. Enhancing the momentum force is the key to accelerate bubble detachment and to minimise bubble size. Recent studies show that replacing steady air with oscillatory air supply could significantly decrease bubble size by enhancing the momentum force. To date, the underlying mechanism for the improved momentum force with oscillatory air supply is not yet clear. This work, for the first time, developed a qualitative model for the momentum force as a function of frequency, on/off time ratio and gas flow rate. It was found that the momentum force was positively correlated with the frequency as well as the air flow rate, and was negatively related to the on/off time ratio. The relation between bubble size and these operating parameters was also investigated. Keywords: Flotation, bubble size, momentum force, modelling, oscillatory air supply

Jan 1, 2020

By P. J. Bartlett, H. A. D. Kirsten

A two-dimensional numerical scheme for the simulation of the plastic stress-strain properties of rock, and the various components of rock reinforcement and shotcrete, was used to rigorously determine the characteristics of tunnel support under squeezing conditions. D-shaped tunnels varying in size from 2,8 to 6,5 m in height and width were considered. The four main mining stages in the BAS block cave on Premier Mine were simulated as alternative sets of field stresses. The rock reinforcement comprised 1,8 m long fully grouted rockbolts and 6,0 m long fully grouted cable anchors at various spacings, and of various tensile and bond strengths. Various thicknesses of shotcrete and footwall support configurations were also considered. It was found that shotcrete and rock reinforcement, owing to inherent differences in stiffness, do not function in unison. As monolithic application, shotcrete has an overriding avidity for the imposed loading and destructs itself unless of extraordinary thickness. Under squeezing conditions, it accordingly supports the excavation surface while the rock reinforcement contains on-going convergence. The loads in the various support components can be determined only by simulation of the shotcrete as a stress-relieved medium to represent its broken condition in the field. The loads in the rock reinforcement vary around a tunnel and are the largest in the sidewalls. The loads increase with an increase in tunnel dimensions, the increase in load in the sidewall due to an increase in height being larger than that in the hanging- and footwalls due to an increase in width. Depending on the strength of the rock, the loads in the rock reinforcement become prohibitively large beyond a certain size of tunnel. The reinforcement of footwall rock can be omitted without adversely affecting that in other walls, provided the tensile and bond strengths of the reinforcement are not limited. The resulting footwall heave may, however, result in undermining of the reinforcement in the adjoining walls. A relative depth of anchor of 1,37 provides adequate bond strength in the larger sizes of tunnel, whereas, in smaller tunnels, anchor length is determined by the allowable convergence. The lack of fun column grouting adversely affects the efficiency of rock reinforcement to a major extent. The loads in rock reinforcement are very sensitive to the strength of the rock. Owing to a lack in refinement and calculational rigour, empirical classification systems are an inappropriate means for the design of tunnel support in squeezing conditions and, as a result, grossly unclear-estimate the support requirements.

Jan 1, 1992

By K. G. Lomberg, M. D. Lionnet

The Bafokeng Rasimone Platinum Mine (BRPM) is a joint venture between Rustenburg Platinum Mines Limited and Royal Bafokeng Resources (Pty) Ltd (Bafokeng). The mining lease has been granted over Boschkoppie 104JQ and is located approximately 10km south of the Pilanesberg Alkaline Complex. The Merensky Reef, which constitutes the principal ore body, is geologically complex. A relationship between its geological variability and the associated mineralisation has been established by defining a number of facies types. The Merensky Reef has been sub-divided into seven distinct facies types: the pegmatoidal facies, contact facies, FW1 pothole facies, FW3pothole facies, FW6 pothole facies, and FW7 pothole facies. The Merensky Reef, which does not form part of the potholed reef, has a large component of footwall mineralisation. The pothole facies, in contrast, has mineralisation located in the hangingwall. Statistical analysis of the facies types indicates that they are unique populations and hence need to be evaluated independently. The mining lease has distinct structural domains that have added additional complications to evaluation of the Merensky Reef. The Merensky facies on BRPM exhibit unique geological and statistical characteristics within well-defined structural blocks. Due to the extreme variability of the Merensky Reef on BRPM a unique approach has been developed, within Anglo Platinum, to evaluate the deposit. Mining cuts are defined and composited for the individual facies types, and then modelled individually by applying known geological controls. The resultant grade models are weighted based on statistical and local geological knowledge. The object of weighting the facies is to replicate the variability observed in the ore body. The introduction of stringent geological controls to the evaluation process has resulted in increased confidence in the in situgrades estimated as an input to the mine planning process

Jan 1, 2006

By G. M. Bentel, G. E. Blight

The environmentally acceptable disposal of fine-particled mining and industrial wastes by the formation of hydraulic- fill slimes dams is becoming an increasingly important aspect of the total mining endeavour. Relatively little is known of the behaviour of waste slurries during deposition. This paper describes and analyses the following aspects of slurry behaviour: (i) the relationship between viscosity, shear strength, and water content, (ii) the slope assumed by a thickened slurry, (iii) particle-size sorting on a hydraulicfill beach, (iv) gradients of hydraulic-fill beaches, and (v) internal erosion during the deposition of slurry. v083n04p073.pdf

Jan 1, 1983

By D. J. Shields

The U.S. Bureau of Mines has developed a microcomputer-based mineral modelling system that enhances analytical capabilities of the agency. It draws upon the extensive data base previously developed for the Bureau's Availability studies. The model is first formulated as a generalized network flow optimization problem, the optimal flows from which are fed into a transportation algorithm. The goal of the modelling system is first to fulfill prespecified demand at the minimum opportunity cost and second to identify feasible, optimal trade patterns. The system facilitates the analysis of policy issues on a real-time basis and supports the Bureau's mission to assess mineral resources.

Jan 1, 1987

By H. W. Glen

The use of a personal style (mainly active voice and first person) makes technical writing more pleasant to read, and helps the writer to express his meaning clearly and to avoid many of the stylistic errors normally encountered in that type of writing.

Jan 1, 1979

By H. Matsuoka, M. Kawata, K. Mitsuhashi, C. Tokoro, T. Kato

Copper sulfide minerals generally have high floatability, but their floatability decreases when surfaces are oxidised under some conditions. Sodium hydrosulfide, NaHS, is known both as an activator and as a depressant for copper minerals. NaHS at low concentrations activates copper oxides by sulfurising surfaces, but at high concentrations depresses copper sulfides by interfering with the adsorption of collectors. In this paper, we applied NaHS to activate oxidised chalcocite and bornite and focused on the conditions of NaHS treatments such as dosages and pH, and the effect of these conditions on the floatability of those minerals. The pure minerals treated with NaHS were analysed using zeta potential analyser and X-ray photoelectron spectroscopy. Analyses indicated that the reactions of NaHS and the products on the mineral surfaces after the reactions (i.e. oxidation, pH change) depended on whether NaHS reacted with copper or iron, and the reactions between NaHS and iron depend on NaHS concentration and pH. Flotation tests using a small column cell were performed with a single pure mineral, and using a Denver-type flotation machine with a copper sulfide ore. The results of each flotation test were utilised to clarify the floatability. The relation between flotation behaviours and analysis results were also discussed. Keywords: Chalcocite, bornite, NaHS activation, flotation, floatability

Jan 1, 2020

By N. Sacks, N. B. S. Magagula, I. Botef

"Low-pressure cold gas dynamic spraying was used to deposit WC-4wt%Ni and WC-4wt%Ni-1wt%Mo coatings onto mild steel. Dense coatings with very low porosities were produced. No decarburization occurred during deposition and no deleterious phases were formed. The coatings were subjected to standardized material characterization tests, as well as slurry abrasion testing to assess their wear behaviour. The wear tests were conducted in synthetic minewater-silica slurries, while distilled watersilica slurries were used as a control. The hardness of the coatings, 512HV0.3 and 458HV0.3 for WC-4wt%Ni and WC-4wt%Ni-1wt%Mo respectively, are comparable to those achieved using high-temperature coating processes. The abrasion wear rates for both coatings were less than 5 mg/min and 10 mg/min in the distilled water-silica and synthetic mine water-silica slurries respectively. The approximately 50% increase in wear rate in the synthetic minewater slurry is attributed to a synergistic corrosive wear mechanism. The predominant wear mechanisms were identified as binder smearing and delamination, with carbide grain fracture and pull-out. IntroductionSlurry abrasion is a common phenomenon in mining operations, particularly in pipelines transporting, for example, sand-water slurries. This type of wear can cause pipeline fracture, shortened operational lifetimes, and impede pipeline efficiency. To reduce or limit abrasion, thermally sprayed tungsten-carbide-based coatings are often used as pipe linings (Kanchan et al., 2010). This is due to their attractive mechanical properties, which include high wear resistance, stiffness, hardness, compressive strength, and tensile strength. These properties are provided by the coating microstructure and composition, which is a blend of hard, wear-resistant tungsten carbide (WC) particles embedded in a fracture-tough metallic binder phase (Brookes, 1997). Cobalt (Co) is the most commonly used binder due to its excellent wetting of the WC phase and moderate mechanical properties; however, its application is limited to non-corrosive environments as it has a low corrosion and oxidation resistance (Andren, 2001). Nickel (Ni) may also be used as a binder, and has been shown to have superior oxidation and corrosion resistance, but in some applications may not have the required mechanical properties (Kim et al., 2006)."

Jan 1, 2016

By H. Ruthert

An understanding of wear in systems conveying solids is extremely important from the standpoints of design and economics. The lifetime of any bulk solids-handling system depends on the type of solids present and the operating parameters of the transportation system. The relationship between wear in backfill pipelines on South African gold mines and the transportation system's operating parameters is well understood. However, the effect of the transported solids on the pipe wear is less well understood, because of an inability to characterize particle sharpness. The sharpness of the particles, together with their size and hardness, is responsible for the aggressivity of particles, or their ability to cause damage and wear. This paper presents a novel method for the characterization of particle sharpness. The results presented rank particles according to the sharpness of their surfaces.

Jan 1, 1993

Council Members 1988/1989, left to right: Row 1 (seated): H.E. James, P.W.J. van Rensburg, Dr H. Wagner, B.C. Alberts, C.E. Flvaz (President), H.G. Mosenthal, R.D. Beck, Professor A.N. Brown, Dr P.R. Jochens Row 2 (standing): P.C. van Aswegen, J.P. Hoftman, D. Wllson, P.M.T. White, K.A. van Gessel, Dr LA. Cramer, J.S. Freer, V.J. Moore, G.A. Brown, R.A. Snodgrass, J.D. Austin Row 3 (standing): Dr G.A. Fourle, Dr J. Lurle, A.M.P. Henderson, R.P. Mohring, J.A. Cruise, J.L. Nel, H. Scott Russell, J.P. Deetlefs, D.A.J. Ross-Watt Inserts, top row: K.R. Beard, N.J. De~ine, R.J. Dlppenaar, G.T.G. Emere, Dr J-P. Franzidls, Professor R.P. King, J.G. McCallum. Bottom row: G.Y. Nisblt, Professor R.P. Plewman, Dr R.E. Roblnson, Dr O.K.H. Steften, P.A. van Wlelllgh

Jan 1, 1989

By C. K. Chabedi

"This paper discusses the difficulties associated with the potential exploitation of the deep multi-seam resources east of the Daarby fault in the Waterberg coalfield. The resources occur at a depth greater than 250 m and the thickness of the coal is roughly 110 m, but the top 50 m comprises coal intercalated with shale and the bottom 60 m contains five seams with sandstone and shale partings. Various factors affecting multiple seam mining at these great depths are discussed with reference to lessons learned from local and international experience on multi-seam mining. Field geological and geotechnical data was utilized to assess the stability of the roof of the seams. There is no specific rock mass rating for the Waterberg area, therefore approximate coal mine roof rating (CMRR) values were used to propose appropriate support strategies. Analysis of Multiple Seam Stability (AMSS) was used to analyse the strength of the parting or interburden between the various seams, the mining sequence, and the interaction between the various seams.The research indicated that it is possible to mine seams with a low CMRR at high mining rates using longwall mining, although support for gateroads is expected to be expensive, time-consuming and onerous to install, and will impact gateroad development rates. It will not be possible to simultaneously mine zones in close proximity and failure of the interburden is predicted, thus dangerous mining conditions are anticipated. However, it will be possible to mine just two of the eleven zones using longwall mining. IntroductionThis paper discusses the difficulties associated with the potential exploitation of the deep multi-seam resources east of the Daarby fault in the Waterberg coalfield using bord and pillar and longwall mining. Figure 1 illustrates the location of the Waterberg coalfield relative to other coalfields in South Africa, while Figure 2 shows the areas containing deep and shallow resources. The Waterberg resources are expected to contribute to South Africa’s future energy requirements, and are currently exploited at the Grootegeluk open pit mine. Grootegeluk produces coal mainly for a power station, with the higher quality product supplied as metallurgical coal. The deposit is technically unique and challenging, being a multi-seam coal deposit with a total of 12 seams over a thickness of 110 m as shown in Figure 3, including Zone 5. There is currently limited knowledge on the multi-seam mining of the deep Waterberg resources. Multi-seam mining utilizing the bord and pillar method has been practised in South Africa before, but at a depth of less than 100 m in the Witbank coalfields. Multi-seam mining in thin seams has also been performed in the Natal coalfields at a depth of less than 160 m, but mostly using bord and pillar mining and secondary mining (partial pillar extraction)."

Jan 1, 2016

By E. J. Maunders

This paper describes briefly a computerised model of the South African electricity system which has been developed jointly by the Atomic Energy Board and the Electricity Supply Commission. The results of an analysis of this system have been subjected to a fairly extensive sensitivity analysis which reveals that up to the year 1990 between 680 and 700 million tons of coal will be required for electric power generation, together with between 7 000 and 10 000 tons of uranium. This result is fairly insensitive to data variations, but the prediction of a cumulative requirement up to the year 2000, of between 1 300 and 1 800 million tons of coal and 25000 to 70000 tons of uranium, proves to be very data-sensitive.