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By H. E. Bartlett

This paper deals with the sampling and mass measurement for ore delivered from a shaft to a processing plant and the contribution of the data from these measurements to the metal balances from shaft deliveries to final metal production. Accurate measurement of the grade and tonnage of run-of-mine ore is important for four main reasons; ? It enables the measurement of the production from different profit centres to within statistically determined confidence limits for daily, monthly, or annual averages. The profit centres could be individual shafts within a mining complex or ore treated on a toll basis ? The monthly production at shaft head is compared to the grades and tonnage determined from underground sampling and mass measurement in terms of a shaft call factor ? The sum of the production from the shafts is the input to concentrators. This input is a major part of the total input into a complex with concentrators, smelters, and refineries. In terms of the Codes of Practice for Metal Accounting the inputs are compared to outputs and inventory changes to assess the efficiencies and unaccounted losses or gains at the various stages in the flow of metal from source to market ? On a daily or daily moving average basis, the grades and tons from shafts are monitored and compared against these quantities from underground measurements. This acts as a control on off-reef mining, dilution, and other factors underground. Accurate measurement of grade of ore at the shaft head has been a challenge because of the large particle size. The conventional wisdom has been that ore can be sampled accurately only after it has been milled to give a slurry that is sampled as feed to a flotation process. However, when the ore fed to a flotation plant comes from multiple sources each source has to be sampled separately. So, relying on the grades determined using the sampling of concentrator input with cross-stream slurry samplers is not an option for determining the grade from an individual shaft. Impala has developed a system for sampling and weighing run-of-mine ore from multiple shafts. The system involves sampling the inputs to the plants using cross-belt (hammer) samplers and weighing the deliveries using in-motion railway weighing systems. Many samples are taken. Individually they have a high variance but, as a consequence of the averaging effect of large numbers and as shown by statistical analysis, the mean results are fit for the purposes of daily grade control for shafts and for monthly accounting of production from individual shafts. In the metal accounting systems at Impala, the measurement of input, including ore and toll-treated material is compared to the output in the form of final metal and tailings losses, as well as any inventory change that takes place, in order to determine a final metal balance from mill feed to product. The calculated inventory is compared to that measured, and the difference, or unaccounted- for metal, is then reported as a percentage of input. The balances obtained over several years yields balances for platinum that are below 1% imbalanced. Keywords metal balancing, platinum mines, AMIRA codes of practice, metal accounting, sampling, run-of-mine ore, statistical analysis.

Jan 1, 2014

By I. M. Macleod

An adaptive steady-state on-line optimizing regulator is proposed as an integrated approach to the control of autogenous run-of-mine milling circuits. This optimizing regulator continuously estimates an internal model of the controlled process, and, based on this, determines mill fresh solids feed rate, mill water flow rate and sump water addition flow rate such that a circuit performance objective is optimized. The dual -requirement of controlling both the mill pulp load and the load of particles of grinding media size is thus directly and simultaneously addressed. The concept and theory underlying the operation of the adaptive optimizing regulator are discussed.

Jan 1, 1987

By J. E. Dickinson, J. L. Sutherland, J. Zhou, M. J. Cole, K. P. Galvin

Interest in early gangue rejection has grown substantially in the minerals industry to reduce energy and water consumption arising from excessive comminution. Recent developments in coarse-particle flotation have proven successful by promoting a quiescent hydrodynamic environment to limit particlebubble detachment and by helping to convey more buoyant particles to the overflow. Current approaches also benefit from upstream classification to remove the finer particles ahead of the coarse-particle flotation. In this study, coarse-particle flotation was investigated experimentally using a novel hybrid gravity-flotation system called CoarseAIR™ that exploits the hydrodynamic advantages of the Reflux® Classifier. The system consists of a lower fluidised bed housing, with a system of inclined channels above. The inclined channels enhance the segregation and hence rejection of intermediate through to relatively coarse particles containing little or no surface liberation. This approach overcomes the need for pre-classification, while offering strong transport of coarse particles made buoyant through their attachment to air bubbles. The paper explores the strong relationship between the solids transport to the overflow and the input gas flux, showing firstly the potential for significant mass rejection at the relatively low gas fluxes, and robust performance in terms of recovery across the full range of gas fluxes. Keywords: Flotation, inclined channels, coarse particle flotation, Reflux Classifier

Jan 1, 2020

By R. M. G. Machunter

The dynamic behaviour of plant circuits and the available instrumentation determine the control strategies that are appropriate for the duty of the mineral separation plant. Although steady state modelling is appropriate to provide the main requirements of grade and throughput capability of a plant, it is the dynamic characteristics that determine the tolerance of the plant to disturbances and also its controller ability. This paper explores these issues with respect to the additional use of dynamic modelling to achieve these objectives.

Jan 1, 2007

By M. S. Shishvan, J. Sattarvand, R. S. H. Meij

Navigating miners during an evacuation using a smart evacuation technology can significantly decrease the evacuation time of an underground mine in case of an emergency. This paper presents a mathematical programming model to calculate the most efficient escape path for each miner as a critical component of the smart evacuation technology. In this model, the total evacuation distance of the crew is minimised and scenarios with blocked pathways, and stamina categories for the miners are simulated. It was found that all the tested scenarios are technically feasible. Using the feature that filters out blocked pathways has no downsides, as safer routes are calculated, and there is no penalty in the computation time. The paper also discusses the social and technical issues that need to be overcome before the algorithm can be implemented as an actual escape solution.

Sep 1, 2021

By G. Blight

As part of an investigation of possible water-saving measures at Anglo American Platinum's Mogalakwena Mine in Limpopo, South Africa, a study was made of conditions on the wet beaches of two adjacent tailings dams, each serving a separate plant on the mine. For one of the dams, a special catamaran was designed and constructed to enable the full length of wet beaches to be accessed and sampled immediately after a layer of tailings had been deposited. For the other dam, the wet beach was accessed using conveyor belting and scaffold planks. The length of the beaches was 400 m on the first dam, and 200 m on the second, and each was sampled at 50 m intervals. The parameters reported in this paper, for both beaches, are: > Water content profiles along each beach from the surface to as deep as 2.5 m, at times ranging from 12 hours to 6 weeks from the end of deposition > Particle size distributions at the surface, down the full length of the beaches > Measurements of feed relative density, tonnage and volume, decant and drain flow volumes, as well as rainfall and evaporation for both dams. The surface water contents shortly after deposition have been used to establish the relative density of the settled, decanted tailings and its variation down the beach. This data, combined with the decant and drainage flows, is used to establish the calculated and actual water recoveries from both dams.

Jan 1, 2012

By J. S. Contreras, E. Córdova, I. Gottreux, A. Anani, A. Ferrada

Cave mining is an underground mass mining technique. The largest projects, which are known as ‘super caves’, produce hundreds of thousands of tons of ore per day, which involves large footprints with considerable column height, and have a life of mine of over 20-40 years. These operations are typically located deep, under high stresses and in competent rock masses, making initiation and propagation of the caving process harder to manage. These challenges must be confronted by optimizing the fragmentation of the orebody to achieve smaller size blocks that will result in consistent caving and improved flow of the ore from the drawpoints. To achieve better performance from the drawpoints, preconditioning is applied to fragment and damage the material required to cave. We present a proposed design for preconditioning in underground mines, considering the challenges that these large-scale mines are already facing, based on a comprehensive analysis of current design parameters, case studies, and sensitivity analyses using numerical models.

Feb 2, 2021

By D. K. Weatherley, T. P. Oladele, L. B. Bbosa

Comminution is a critical stage of mineral processing which aims to reduce the size of ore particles through breakage, consequently increasing the likelihood of liberation of valuable minerals. Several factors, such as pre-existing cracks, mineralogical composition and particle size, are known to affect ore breakage behaviour. While investigating the contributing role of each factor is essential to improve understanding of breakage, isolating individual factors by experiment is typically impractical. Numerical techniques such as the Bonded Particle Model-Discrete Element Method (BPM-DEM) have been developed as a means of studying closely controlled breakage conditions. This study evaluates this technique in modelling the effect of pre-existing cracks breakage behaviour for synthetic ore specimens during impact breakage in a short impact load cell (SILC). Synthetic ore was modelled as a collection of discrete entities connected by brittle-elastic beams. Model parameters were calibrated against macroscopic material properties of a conventional homogenous synthetic ore. The extent of pre-existing cracks within specimens was prescribed by varying discontinuities between the connecting entities. An approach with the insertion of seed points was used to model the mineralogy of the specimens. Results showed that pre-existing cracks resulted in a decrease in the fracture force as the crack density increased. Additionally, the irregularity of fracture patterns within rock specimens after impact became more apparent as the crack density increased. The findings were used to emphasise some of the attributable causes of the variation observed between ore hardness characterisation tests and in situ ore behaviour. Keywords: Element Method; Bonded Particle Model, synthetic rock sample; rock fracture, preexisting cracks

Jan 1, 2020

A modified germanium recovery process in conventional zinc metallurgy, based on the understanding that ferric hydroxide has strong ability to coprecipitate germanium element, was presented. The relaxed lab-scale experiments were conducted to evaluate its technological feasibility. The experimental results obtained show the new process has great competitive advantage over the conventional one.

Jan 1, 2008

By P. J. Petit

Changing mining conditions, legislative issues, rising costs, and the constrained supply of skills and electricity in South Africa, have been some of the criteria forcing mine operators and engineers to review the traditional use of compressed air to power underground hand-held drilling operations in narrow-stope, hard-rock mines. In answering the question, 'What is the best way to power this mine?', they have resorted to trade-off studies with results of limited usefulness, because they are site-specific, and use only a few comparative parameters, based on a small data set. This paper aims to provide definitive, objective and quantitative decision-making approaches to evaluate alternative energy delivery systems for stoping in narrow-reef hard rock mines, based on primary and relevant criteria. These include quantifiable criteria such as capital expenditure, operating costs, resulting financial returns, energy and water usage, production performance, life-of-mine, etc. Other important criteria with qualitative aspects are integrated. These are safety, health and environmental concerns, legislative requirements, business risk, technology change, etc. The paper begins with a review of decision making processes available, and proposes the use of the analytical hierarchy process (AHP) method as a multiple-criteria decision making tool and the decision tree-based on milestone-driven uncertainties, to validate the selection of the energy delivery system. The results of a multiple-criteria questionnaire completed by a variety of industry experts and professionals, such as production and operation managers, consultants, suppliers and techno-financial analysts, are discussed. The evaluation indicates the potential for hydropower to be the best solution for narrow reef hard rock mines, based on current information, mine design layout and production.

Jan 1, 2012

By W. J. C. Pothas

This paper describes the state of the art in backfilling techniques used on East Driefontein Gold Mine. Areas covered are the preparation plant, the distribution plant, the pipeline network and finally placement techniques. The in-situ performance of placed backfill has been assessed by the Chamber of Mines Research Organization, and results to date will be shared. Details of the backfill expansion programme that will enable this facility to be made available to all working places are included. Benefits obtained by the introduction of backfilling are described, together with certain management aspects to be considered in the control of quality and quantity of support supplied via a pipeline.

Jan 1, 1988

By Q. G. Reynolds

The dual-electrode DC arc furnace, an alternative design using an anode and cathode electrode instead of a hearth anode, was studied at small scale using computational modelling methods. Particular attention was paid to the effect of two key design variables, the arc length and the electrode separation, on the furnace behaviour. It was found that reducing the arc length to brush arc conditions was a valid means of overcoming several of the limitations of the dual-electrode design, namely high voltages and arc deflection.

Jan 1, 2012

By D. Amariei, G. Bartolacci

Preconcentration of precious and base metal ore types is becoming more relevant with the discovery, the development and the operation of many lower grade deposits around the world. Ore preconcentration designates an operation consisting of sorting an ore to reduce the quantity of waste material sent to the concentrator. It can be seen as rejecting waste material as much as possible, without compromising the recovery, or as the recovery of as much valuable metal as possible into an optimum mass yield. Benefits of preconcentration are multiple: grinding energy savings, costs reduction, lower environmental footprint, extension of the reserves, etc. However, all of these benefits do not apply simultaneously and are not applicable to each ore type. Technologies employed for preconcentration are highly dependent on ore properties, especially on the particle size at which the gangue is liberated. Thus, from coarser to finer particle sizes, technologies such as screening, ore sorting, gravity separation (Dense Media Separation (DMS) and jigs) and coarse flotation could be used for the rejection of liberated gangue. The difficulty starts when a benchmark is needed to select a suitable technology, especially for low-grade gold deposits. This paper presents the challenges of comparing lab-scale technologies (DMS versus jig) for gold preconcentration. The experimental data will show to what extent comparing two lab-scale gravity technologies for gold preconcentration can be challenging. An attempt to evaluate the energy savings by assessing the hardness of the tailings after gravity separation will also be presented. Keywords: Challenges in gold preconcentration, gravity separation, jig, dense medium separation

Jan 1, 2020

By A. J. Barrett

Backfilling at Vaal Reefs during the experimental phase has been carried out using cyclone-separated tailings which are transported underground hydraulically by means of a borehole. The selected tailings are piped from the borehole to the stope and then, after the addition of a flocculant, deposited in place via single or multiple nozzles. The successful performance of the backfill/stope system relies not only on the actual quality of the placed material, but also on the effective operation of stope filling. The success of stope filling relates to the complex, time- dependent interaction of backfill, stope and barricade materials, and the transport and placement procedures used. A comprehensive laboratory testing programme was carried out on backfill materials at Vaal Reefs in conjunction with the experimental stope- filling operation. The range of testing techniques corresponds to the requirements of each stage of the backfilling process. Where necessary, new or alternative techniques were sought during the course of the experiment to satisfy and facilitate better understanding of the different processes. These techniques comprise both routine quality control testing and research-oriented testing. The laboratory testing techniques appropriate to each stage of the backfill process are briefly discussed. This paper presents the main conclusions of the study and identifies areas for further research.

Jan 1, 1988

"Chairman: Professor E. P. PFLEIDER Rapporteur: Mr P. R. JANISCHPapers:Optimum design of an open pit by D. A. PhillipsAn open pit design system for stratiform ore bodies by L. A. J. Pronk van Hoogeveen, J. R. Cutland and M. WeirAn open pit design model by R. H. Robinson and N. B. PrennThe Chairman in opening the session highlighted the current situation in the United States of America where some 85 to 90 per cent of all mining (except for coal) is done by open pit methods, and where some 40 per cent of the coal is mined by open pit. The South African industry is, of course, dominated by the deeper level underground gold mines; however, with the aid of an extensive open-pit technology developed in other countries, South African companies arc continuing to explore possibilities of open-pit exploitation of base mineral, coal and diamond deposits. Several successful ventures are already in operation at Phalabora, Optimum Colliery, Kimberley and elsewhere, and South Africa is undoubtedly benefitting in this regard by drawing on the experience of others.The papers of this session were therefore particularly pertinent, particularly when viewed in conjunction with those presented at an earlier South African conference devoted specifically to open pit mining.Two themes dominated the discussion which followed presentation of the three papers at this session. The first related to the practicability of using the reported techniques as a tool for management, and their acceptability as such to mine management. This was, of course, not peculiar to this particular session, but rather the major point of debate over the entire conference.The second theme concerned the reliability of the initial drill hole data on which any open-pit design must necessarily be based. This gave added emphasis to the problem of practicability, for no management is likely to accept an elaborate pit design unless confidence can be placed in the ore reserve assessment.In noting these two features of the discussion, it should be remembered that all the three papers dealt largely with design and pre-planning; there was an unfortunate lack of papers or discussion on the subject of controlling the day-to-day mining of open pits, and this is an area where the potential impact which can be made by the computer is even greater.In introducing his paper, Mr D. A. Phillips pointed out that it was concerned only with optimization, and that it therefore assumed, firstly, an adequate subdivision of the orebody into ore blocks characterized by grade, rock type, structure, etc., and, secondly, an acceptable forecast of costs and mineral prices in order to generate a suitable profit matrix. However, programs dealing with these problems for both massive, disseminated and stratiform ore bodies are available to Roan Consolidated Mines. The method of optimization which he had described had been used on a region split into 28 000 blocks, for which the program used three hours of C.P.U. time on an IBM 360/40 computer. The computer design had provided a three per cent improve¬ment on a normal design of the same area."

Jan 1, 2014

Outotec?s direct leaching process is an atmospheric leaching process for sulphidic zinc concentrates. The direct leaching process, based on special leaching reactors developed by Outokumpu, is the most economical way to produce electrolytic zinc. The production scale applications of the process realized so far are in use at Boliden?s Kokkola and Odda zinc plants. A new full-scale application of Outotec?s direct leaching process will be started in 2008 in China. The process is integrated into the existing production plant, replacing part of the old production facilities by modern technology. Besides high zinc extraction the plant expansion and modernization offers also indium recovery, iron precipitation and solution purification processes. The process equipment includes tailor-made reactors designed on the basis of the work and experience of Outotec research centre. Keywords: zinc, concentrate, atmospheric leaching, technology, indium, iron precipitation, solution purification, Outotec.

Jan 1, 2008

By J. L. Louw

The utilization of centralized blasting systems is progressively becoming a feature of the South African mining industry, owing to increased pressure from the Department of Mineral Resources (DMR) to ensure safer blasting conditions. The purpose of a centralized blasting system (CBS) is to ensure reduced blasting-related incidents and to improve the underground working environment for the mine?s employees. The brief from Samancor Western Chrome Mines (WCM) pertaining to this study was to explore?at a conceptual level?all variables pertaining to the implementation of the CBS. Studies have indicated that Sasol?s SafeBlast? was preferred due to low overall risks and user-friendliness. Empirical observations, however, indicated that attention should be given to the current shaft-clearing procedure and to the installation of additional transformers. This is required to co-ordinate the power supply from 2 level to 10 level with the required supply of the network slave station. A cost analysis indicated that an electronically initiated CBS will result in a R24 cost increase per panel, which relates to a mere 1% cost increase per panel. This cost, in addition to the R1.1 million capital cost, proved to be negligible in comparison to the possible direct and indirect costs incurred by the mine in the event of a blasting-related incident. It was found that Samancor should conduct further studies incorporating the necessary requirements to ensure a successful transition, while considering all stakeholders on the mine. This study proved that the transition to a CBS would ensure safe and efficient blasting at Samancor?s Mooinooi Mine.

Jan 1, 2012

By J. O. Roux

The Impala Platinum Limited Base Metals Refinery in Springs, South Africa, refines convertor matte containing base metals and PGM?s (platinum group metals) using a hydrometallurgical process. In addition to a PGM-rich concentrate that is further refined, the Base Metals Refinery produces nickel (in powder or briquette form), copper cathodes and cobalt powder. Five leaching stages are utilized to remove the base metals and impurities from the PGM concentrate. The project described in this paper involves a radical redesign of the second stage leach autoclave with the aim of reducing maintenance costs and downtime, increasing throughput and ensuring a safer working environment. The second stage leach process at the Base Metals Refinery, produces a PGM-rich residue through the extraction of copper and any remaining base metals from the first stage leach solid residue material. Leaching is performed in an autoclave operating at a pressure of 600 kPaG and a temperature of 140°C to 150°C. Concentrated H2SO4 is added to the feed to produce a solution with a free acid content between 10 and 30 g/L and a pH of less than 2. Oxygen is passed into the 1st and 2nd compartments of the autoclave. As a result of the aggressive, highly oxidizing conditions within the autoclave, the original design utilized a heavy wall carbon steel shell with lead lining and two layers of acid bricks. The bricks acted as a wear resistant material and ensured a low surface temperature at the skin of the lead lining, whereas the lead lining acted as a corrosion barrier to protect the steel shell. These brick lined autoclaves were heavy, maintenance intensive and prone to catastrophic failure. A radical autoclave redesign was needed to reduce the high costs of manufacturing, installation and operation. Finite element modelling was used to optimize the design and various high alloy materials were investigated for the application. Based on extensive test work, the carbon steel shell was replaced with duplex SAF 2205 stainless steel in the new design. The use of this highly corrosion resistant material, eliminates the need for the lead and brick linings and reduces the wall thickness requirements substantially, resulting in an increase in the available volume for the same external shell dimensions. The vessel is welded in circular sections to form an elongated, impermeable, cylindrical body, with several agitator nozzles and compartments. The first new duplex stainless steel unit has been in continuous operation for more than 24 months and its performance has exceeded the highest expectations. The design was so successful that a patent has been registered. The major benefits of the new stainless steel design include increased throughput, improved performance, less downtime and a safer working environment. Lead has been eliminated and replaced with a more environmentally safe material, and the total lifespan of the vessel has been extended. In addition, the autoclave project has made significant developments in advancing the field of welding and optimizing the integrity of welding quality standards in South Africa.

Jan 1, 2009

By B. P. Watson

The Bushveld platinum group metal deposits in South Africa are the largest in the world. These deposits occur as two distinct stratiform tabular ore bodies and strike for many hundreds of kilometres. Mining is extensive, with depths ranging from close-to-surface to 2300 m. The mining method is a variation of planar open stoping. Crush pillars are widely used to support the open stopes. These pillars are designed to fail and the residual strength provides the required support resistance to stabilize the stoping excavations. This paper describes the in situ measurement, of stress within a Merensky pillar from Impala Platinum. These measurements were used to derive a stress-strain curve that includes pre and post failure behaviour. 2D FLAC modelling, with strain softening, was done to investigate how damage expands into the hanging- and/or footwall. Although the so-called ?squat? effect is always present, the modelling suggests that it does not dominate the pillar system behaviour at larger width to height ratios. Punching of the pillar into the hanging- and/or footwall increasingly controls the pillar system behaviour with larger width to height ratios. The paper investigates the behaviour of a system that incorporates the immediate hanging- and footwall, as well as the pillar itself.

Jan 1, 2009

1. Recent publications Canadian minerals yearbook 1986. Review and outlook. Mineral Report 35. Available from Printing and Publishing, Supply and Services Canada, Hull, Quebec KIAOS9, Canada. $44.70. The 1987 Sigma cataloge. SI. Louis (USA), Sigma Chemical Co., 1986. 1736 pp. Der Maschinenschaden. Munchen(G. rmany), Koningstrasse 28. Metallurgicalworks in Canada primary iron and steel 1987. Mineral Bulletin MR 214. Available from the above address at $6.60 per copy. Annual report 1986. Oxbridge, International Research Institute, 1987. Canadian iron ore industry statistics, 1986, by B.W. Boyd and A. Cadieux. Publication MRI 87/6. Micromedia Limited (Document Delivery, 165 Hotel de Ville, Place du Portage 2, Hull, Quebec J8X 3X2, Canada). Tin and its uses no. 153. Oxbridge (England), International Tin Research Institute, 1987. This issue includes articles on can-making, titanium-tin alloy, coppernickel2. New journal Pergamon are producing a new journal, entitled Minerals Engineering, which they describe as 'an international journal devoted to innovation and developments in mineral processing and extractive metallurgy'. 3. Mintek publications Report M298. The effect of flocculant additions to the leaching solution on the kinetics of the oxidative dissolution of sphalerite, by F.K. Crundwell. Report M304. The smelting of ilmenite in a d.c. transferred-arcplasma furnace with a molten-anode configuration, by AD. Brent. Jut 1987. 17 pp. Report M312. The reactions of some thiol collectors at noble-metal and pyrite electrodes, by D.R. Groot. Jut 1987. 21 pp. Report M321. The determination of molybdenum, arsenic, and uranium in activated carbon by X-rayfluorescence spectrometry,by A.M.E. Balaes and G.J. Wall. Aug. 1987. Report M321. The determination of molybdenum, arsenic, and uranium in activated carbon by X-rayfluorescence pectrometry, by A.M.E. Balaes and G.J. Wall. Aug. 1987.

Jan 1, 1988