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By J. C. M. De Souza

The Samarco pipeline, the longest in the world for iron ore slurry transportation, started operation in 1977 to transport concentrate from the Germano facilities, in the state of Minas Gerais, to the pelletizing plant, located 396 km away, in the state of Espirito Santo, on the south-east coast of Brazil. Designed for a nominal capacity of 12 million tons per year, the pipeline reached 15.5 million/year due to operational improvement and, in late 2005, it was resized to transport 16.5 million tons per year, with the installation of a booster station. In order to ensure the accurate performance of the system, a pipeline cleaning programme was carried out and, since it has been 29 years from the operation start-up, precise knowledge of the pipe condition would be fundamental to guarantee the success of the project. The cleaning programme demonstrated the potential of operational gains from this initiative, and its effect on the corrosion control and future measures, which should be permanently adopted to maintain the improved capacity. The intelligent pig runs provided essential data used to establish a pipeline integrity programme that allowed Samarco to develop a pipe repair plan based on the severity and significance of the features and to implement a future pipe replacement plan.

Jan 1, 2007

By O. G. Motlogelwa

Any company operating a fleet of trucks will be greatly advantaged if the associated fuel costs can be decreased. The optimal tonnage that trucks should be hauling to ensure minimum consumption of diesel fuel has been investigated. The fuel consumption in litres per ton for a fleet of trucks hauling different tonnages of iron ore was recorded in a time-motion study and the optimal tonnage was determined. The following points emerged as a result of the research: > The diesel fuel consumption in litres per ton decreased with increasing tonnage, which means that in order to improve fuel efficiency, the trucks should carry 21 t > As a result of queuing, Africa Truck Hire (one of the two equipment contractors) loses R282.57 per day to diesel that the trucks burn when they are left idling. It is thus advisable that the company adopt a policy of requiring the drivers to switch their engines off while queuing at the stockpile > Filling up the trucks during each shift or rationing the quantity of diesel to 200 litres each shift has a negligible impact on the trucks' fuel efficiency > For the company to spend 35% of their revenue on fuel, the diesel price would have to be between R8.14 and R8.78 per litre.

Jan 1, 2013

A set of six design principles for rock engineering was developed by Bieniawski: design objectives; minimum uncertainty; simplicity of design components; state-of-the-art practice; optimization; and constructability. These principles translate into a ten-step design process. Although this methodology was developed for rock engineering, it is applicable to any design process, feasibility study, etc. Diligent use of such a process will ensure that a defensible design or evaluation has been carried out. Recently a ?strategic conversation? process has been identified by Ilbury and Sunter that includes the following ten steps: scope; players; rules of the game; key uncertainties; scenarios; SWOT analysis; options; decisions; measurable outcomes; and meaning of winning. A comparison of these steps with those of the design process shows a remarkable correlation. This is perhaps not surprising since engineering design is a strategic up-front issue, and front end loading in a project is critical to achieve the expected project performance. Based on the close correlation between the strategic conversation and engineering design processes, a circular design process, rather than a linear process, is proposed. The ?circle of design? better reflects the importance of review and monitoring in the two phases of successful design and implementation?defining the design, and executing the design.

Jan 1, 2009

By P. J. Foster

This paper summarizes a behavioural safety project undertaken in seven quarries in the UK in order to investigate its applicability as a means of raising the standards of health and safety in the UK quarrying industry. A behavioural safety methodology was developed and trialed at two quarry sites in the UK. During the study the number of unsafe acts and minor accidents had decreased and at one of the quarries there was a 95% reduction in the number of working days lost. It was then introduced at a further five sites of varying sizes and within this focused specifically on lone workers as well as behaviours that affected quality and environmental issues. Positive improvements were noted in most of these operations. The project also looked at the effectiveness of health and safety performance indicators. Here, the common lead and trailing indicators for both safety and health that are commonly used by industry were identified. In consultation with the UK quarrying industry these were then used to identify specific performance measurement criteria that can be used by the quarrying sector to measure performance. Following on from this, a quarry specific health and safety management system specification was developed for large and small quarries.

Jan 1, 2008

The development of short- and medium-term mine production schedules in isolation from each other has meant that only a local optimum can be achieved when each scheduling phase is carried out. The globally optimal solution, however, can be achieved when integrating scheduling phases and accounting for the interaction between short-term and medium-term activities simultaneously. This paper addresses the task of integrating short- and medium-term production plans by combining the short-term objective of minimizing deviation from targeted mill feed grade with the medium-term objective of maximizing net present value (NPV) into a single mathematical optimization model. A conceptual sublevel stoping operation comprising 30 stopes is used for trialling segregated and integrated scheduling approaches. Segregated medium- and short-term scheduling using separate models achieved an NPV of $42 654 456. The final scheduling approach involved integrating the two scheduling horizons using the newly-developed globally optimal integrated production scheduling model to achieve an NPV of $42 823 657 with smoother mill feed grade. The larger the stope data set, the larger the difference between the two scheduling approaches is likely to be. At the very least, an integrated approach ensures feasibility across the two scheduling horizons, which cannot always be assumed when using a segregated approach.

Jan 1, 2012

By J. C. M. De Souza

The Samarco pipeline, the longest in the world for iron ore slurry transportation, started operation in 1977 to transport concentrate from the Germano facilities, in the state of Minas Gerais, to the pelletizing plant, located 396 km away, in the state of Espirito Santo, on the south-east coast of Brazil. Designed for a nominal capacity of 12 million tons per year, the pipeline reached 15.5 million/year due to operational improvement and, in late 2005, it was resized to transport 16.5 million tons per year, with the installation of a booster station. In order to ensure the accurate performance of the system, a pipeline cleaning programme was carried out and, since it has been29 years from the operation start-up, precise knowledge of the pipe condition would be fundamental to guarantee the success of the project. The cleaning programme demonstrated the potential of operational gains from this initiative, and its effect on the corrosion control and future measures, which should be permanently adopted to maintain the improved capacity. The intelligent pig runs provided essential data used to establish a pipeline integrity programme that allowed Samarco to develop a pipe repair plan based on the severity and significance of the features and to implement a future pipe replacement plan.

Jan 1, 2007

By L. Capponi

At Tapira Mine (Tapira/MG/ Brazil), the largest phosphate concentrate producer in Brazil, the mining is fully mechanized and carried out using open pit mining. This work presents a review of the sampling protocol used for short term mine planning at different stages. Sampling is carried out in two steps. Firstly, diamond drill cores are sampled at 50 x50 x5 metre pattern. Fosfertil analyses grades of P2O5, Fe2O3, CaO, MgO, SiO2 and TiO2 plus geological logging. This information feeds the long term block model. Secondly, drill cuttings are collected at the mining faces released for blasting. The drilling pattern is staggered and when blasting is not necessary, a drill hole is made centered in the block area (25 x 25 x 5 m). The primary samples have mass approximately 60 kg and the same chemical parameters are analysed. Metallurgical and mass recovery tests are performed as well. These samples are used to build the short term block model. At the final stage, the short term mine planning model is used to build a stockpile blending model which is simulated using a spreadsheet to match the goals established by the process plant.

Jan 1, 2009

By Z-B. Cheng, X-J. Liu

"Intensive mining of shallow coal seams tends to result in severe subsidence of the ground surface. Surface subsidence is a complex dynamic spatiotemporal process that constantly changes as mining advances. With Liangshuijing coal mine as a case study, 3DEC numerical software was used to simulate the development of the moving subsidence field from open-off cut to full subsidence on the working face. The results demonstrate that the formation of a subsidence field is directly related to the coal seam depth H and the extent of the mine workings. Taking the advance distance of the working face as the unit of scale, the process of surface subsidence can be divided into four stages: micro-change (H/ p), the development of surface subsidence (2H/ p), the formation of a movement subsidence field (3H/ p), and dynamic balance (4H/ p). The changing movement of a subsidence field can be fully described by the characteristics of surface subsidence value, curve slope, and the inflection position. The results provide important technical support for predicting surface subsidence and the temporal–spatial evolution features of mining subsidence. IntroductionChina’s coal output was about 3.52 Gt in 2017 (China National Coal Association, 2018), making it the largest coal producer in the world, accounting for almost half of total world coal production. Coal production and distribution play a vital role in the energy industry of China. However, entire overlying strata, from coal seam to surface, are disturbed as a result of long-term, high-intensity and large-scale mining, causing surface subsidence. This results in groundwater loss, surface desertification, vegetation loss, and mud-rock flows. These issues have become topics of common concern throughout society. According to related research results (Liu 2008), the area of surface subsidence caused by coal mining in China has reached about 600 000 ha, some 0.2 ha (and up to 0.42 ha) of surface subsidence for every 10 000 t of coal produced. Surface subsidence disasters not only damage buildings, water conservancy, transportation infrastructure, and farmland, but also cause adverse effects on individual and community lifestyles, environmental hygiene, and economic development. Deterioration of the ecological environment in mining areas has a long history and is very harmful (Lu, 2015)."

Feb 1, 2019

By R. W. Ottermann, N. D. L. Burger, A. J. von Wielligh, P. R. de Wet, J. L. Steyn

Underwind and overwind protection system concepts for mine hoist shafts were developed in conjunction with and for the Safety in Mines Research Advisory Committee (SIMRAC). End-of-wind operation of mine hoists is the most hazardous aspect of mine hoisting, carrying the highest risk of loss of life, injuries to workers and loss of production. Systems operating mechanically and with local actuation were developed. This was done in order to remove the increased risk associated with dependency on remote or external energy and information supply associated with most existing protection systems. The proposed underwind protection system concept absorbs the energy of motion of a conveyance overrunning the design lower limit of travel by drawing a metal strip through a set of rollers. This action causes dynamic cyclic plastic bending of the strip material that converts the kinetic and potential energy of the conveyance into strain energy of the metal. In the concept design a pair of cable slings attached to the strips catch the overrunning conveyance and transfer the retardation force. The proposed overwind protection system concept absorbs the energy of motion of a conveyance overrunning the design upper limit of travel by early detaching of the conveyance from the hoist cable. This detaching is carried out via an additional humble hook activation mechanism fitted at a sufficient height below the spectacle plate to allow the conveyance to retard to standstill under gravity before it would crash into the spectacle plate. The conveyance so brought to rest is prevented from falling by means of jack catches on the conveyance interacting with a rack (toothed profile) fitted on the guide rails in the retardation zone. Scale models (1:10 scale) of both protection systems were designed for conformance with established retardation standards. The retardation standards were limited by the requirement of passenger safety in retarding cages. The models were then built and tested in a 1:10 scale shaft model. Retardation performance close to the required levels was achieved. The retardation distance required should allow such systems to be retrofitted to most existing shafts. Given the simplicity and robustness of the designs

Jan 1, 2003

By T. Oberthür

The Great Dyke of Zimbabwe constitutes the world&apos;s second largest reserve of platinum group elements (PGE) after the Bushveld Complex in neighbouring South Africa. Within the Great Dyke, economic concentrations of PGE are restricted to sulphide disseminations of the Main Sulphide Zone (MSZ), which are currently mined at the Ngezi, Unki, and Mimosa mines. Near-surface oxidized MSZ ores have a large potential. Their total resources are in the range of 160-250 Mt; however, all previous attempts to extract the PGE from this ore type have proved uneconomic due to low PGE recoveries (<< 50 per cent) achieved by conventional metallurgical methods. Within the ores of pristine, sulphide-bearing MSZ, the PGE are bimodally distributed. Platinum occurs mainly in the form of discrete platinum group mineral (PGM) grains (mainly bismuthotellurides, sulphides, and arsenides), whereas approximately 80 per cent of the Pd (and some Rh) is hosted in pentlandite. Within the oxidized MSZ ores, the PGE are polymodally distributed. Whereas the arsenide- and sulphide-PGMs that make up approximately 25 per cent of the original Pt content of the ore largely remain stable (relict PGMs), the remaining PGMs are disintegrated. The base metal sulphides are destroyed, partly releasing their base metal and PGE contents, and are replaced by iron oxides or hydroxides. Unspecified amounts of the PGE are redistributed and either form secondary PGMs, are found in chemically and mineralogically ill-defined (Pt/Pd) -oxides or hydroxides, or in iron-hydroxides, Mn-Co-hydroxides, and in secondary silicates. The problematic processing of oxidized MSZ ores is attributable to their complex nature and polymodal distribution of the PGE, prohibiting a significant upgrading of the ores by conventional metallurgical methods. Therefore, only bulk leaching methods are viable for ore treatment, and novel metallurgical methods have to be developed for the processing of these ores. Our ongoing work aims at locating the PGE in their mineralogical form in order to understand the mineralogical balance of the PGE in the ores and thereby facilitate the evaluation of metallurgical options for their recovery. A short overview on options and recent advances regarding the recovery of the PGE from oxidized ores is given.

Jan 1, 2013

By C. Klutke

This paper outlines a simple methodology for mineral characterization, developed as part of the Australian Mineral Industry Research Association (AMIRA) managed research project P777 ?The Development of Heavy Suspension Techniques for High Density Separations (Replacement of Clerici?s Solution)?. The project was sponsored by De Beers, Rio Tinto and Iluka Resources. Heavy mineral characterization of samples arising from exploration, mining or metallurgical processes is frequently conducted using laboratory heavy liquid analysis. Unfortunately, there are only a limited number of high density (?heavy?) liquids and these tend to be more toxic as their density increases. Low-toxicity inorganic solutions, based on tungsten compounds, have been developed that can be utilized at relative densities (RD) up to 3.0. Beyond this value organic liquids can be used; however, this presents significant health and safety hazards. Diiodomethane (methylene iodide) having a relative density of 3.31 is commonly used. Mixtures of thallium formate and thallium malonate were found in the early 1900s by Clerici to provide liquids having specific gravities between 4.0 and 5.0. For the characterization of the heavy components of mineral sand deposits (e.g. anatase RD 3.9, rutile RD 4.2, ilmenite RD 4.4?4.7 and zircon RD4.6?4.8) there is currently no heavy liquid alternative to Clerici?s solution. Clerici?s solution is highly toxic and testing is now conducted by few laboratories worldwide, with costs reflecting the chemical costs, infrastructure costs and health and safety regimes (e.g. blood testing of exposed staff). A simple laboratory technique of density fractionation has been developed, employing suspensions of fine tungsten carbide particles in lithium heteropolytungstates solutions, that can replace Clerici?s solution in the evaluation of fine mineral sands samples (e.g. -250 +150microns). The developing methodology that can achieve low-cost, low-toxic separations at relative densities above 4.0 is outlined and the comparison of results with Clerici?s solution presented. In addition, preliminary work on density fractionation of zircon samples is presented. Zircon fractionation relates to their inclusion, radionuclide content and metamictization.

Jan 1, 2007

This overview serves as an introduction to the succeeding paper in this issue, that entitled &apos;The design of backfill as a support for Polish coal mines&apos; by Professor J. Palarski. As such, it briefly describes the coal-mining methods, production, and mining education in Poland. It ends with a brief account of Professor Palarski&apos;s qualifications and experience in the fields of backfilling and hydro transport.

Jan 1, 1994

Mr Philipus Albertus von Wieligh, a former President of the Chamber of Mines and Executive Director of Johannesburg Consolidated Investment Company, died in Johannesburg on Thursday, 28th December, 1989. He is survived by his widow, a daughter, and two grandchildren. Mr Von Wieligh served the mining industry for 45 years. The P.A. von Wieligh Park in Hilshaven on the West Rand is named after him.

Jan 1, 1990

By J. B. Parry

Three types of high-strength cold-rolled steels are discussed. Two are currently being developed at Iscor: rephos phorized steel and niobium microalloyed steel. Heat treatment after cold rolling is carried out in batch annealing furnaces. A third type, dual-phase steel, which requires a continuous annealing treatment for economic production, is also described. The properties, strengthening mechanisms, and applications of these three steels are examined.

Jan 1, 1983

By T. V. Thoms, T. R. Twidle, J. W. Crous

This note describes an investigation into the most cost-effective material for mil-shell liners in large-diameter pebble mils with lifter bars. This is part of a project being conducted at the gold plant of East Driefontein near Carletonvile in the Transvaal. It was found that manganese-steel and chromium-steel liners are more cost-effective than liners of cast mild steel or white iron.

Jan 1, 1990

By J. Palarski

Long waling is used in most Polish coal seams, with backfilling to support the roof, to create an artificial hanging wall to prevent movement if the rockmass, and to eliminate the fire hazard in abandoned mine workings. This paper discusses the results if tests conducted on 7 different backfill mixtures, and gives equations for the estimation if the flow if backfill slurries through gravitational pipelines.

Jan 1, 1994

By A. C. Wingrove

The paper outlines the selection of a dewatered tailings backfill system for the Johannesburg Consolidated Investment Gold Division, and covers the following aspects: (i) 16 Shaft pilot testwork; (ii) engineering design problems relating to the Cooke 3 Shaft backfill operation; (iii) in-line mixing - applications with respect to the Randfontein Estates Gold Mine and Western Areas Gold Mine backfill operations; and (iv) pipeline friction losses.

Jan 1, 1988

By E. B. Viljoen

I. R. M. Chaston* As a gravity concentration man in the midst of so many eminent flotation experts I feel like Daniel. However, I would respectifully disagree with the opening remarks by Mr Viljoen. First, he suggested that gravity concentration of cassiterite was not effective in sizes below 30 microns. In a paper to the Institution of Mining and Metallurgy in 19621, I gave some figures for gravity concentration of deslimed material both for cassiterite and for wolfram which indicated that careful shaking table concentration of fine deslimed material could give recoveries of 75 to 90 percent in the size ranges above 13 micron. These were figures obtained from normal plant operation. It would be interesting to know if any tests were carried out on gravity concentration of the efficiently deslimed feed to this flotation plant to see what the recovery would be. Secondly, Mr Viljoen suggested that gravity concentration plants were expensive. He gives the cost of a flotation plant at R300 000 to treat a feed of 400 t.p.d. to the desliming section or 160 t.p.d. to the flotation cells. From his operating cost figures it would appear that the flotation cost is approximately R2 to R4 per ton of feed to flotation and the total operating cost about R2 per ton of feed. A table plant to treat this tonnage of deslimed feed would consist of about 30 tables and should cost considerably less to install and very much less to operate than the flotation plant. Third, Mr. Viljoen suggests that the concentrate grade from a gravity concentration operation would be unacceptably low. The paper quoted abovel showed that once a primary fine concentrate had been made, giving a suitably deslimed material, further gravity concentration could give concentrates of over 40 per cent Sn with high recoveries. Tailing from this dressing stage was naturally returned to the feed. To give these high grades, sulphides had been removed by flotation but this necessary flotation step is simple and cheap to operate and could be restricted to the final stage when the quantities involved are small. I would agree that flotation has a part to play in fine tin recovery but this role would only seem to become fully effective at sizes below 10 micron and would seem to require, as Mr. Viljoen has suggested, a mechanism for sizing which is effective down to, say, 1 micron. In the meantime it would be interesting to try tabling the flotation tailing to see if it was possible to recover the plus 20 micron cassiterite which is not recovered well by flotation. One point which has been made during discussion is the suggestion that cassiterite particles have a fully hydrated surface at normal pH. I wonder if this layer could materially affect the overall specific gravity of a very small cassiterite particle and, if so, whether gravity concentration of the very fine cassiterite would be improved if carried out in acid conditions, which would inhibit the formation of this hydrated surface. REFERENCES 1. CHASTON, I. R. M. &apos;Gravity concentration of fine cassiterite&apos;. Trans. Inst. Min. and Met. Vol. 71, p. 4, 1961 and 1962, and discussion of this paper. Dr. R. P. King Mr Viljoen has drawn attention to the difficulties associated with the presence of very fine particles in the flotation pulp and he has indicated that, on the Union Tin plant, they must sacrifice 30 percent of the cassiterite in the minus 7 /Lm material which they separate and lose. He has also indicated that the rate of flotation of cassiterite remained high at these small sizes. The problem arises because the very small particles stabilize the froth making it impossible to handle. These stable froths cannot be easily broken and therefore cannot be pumped and processed. It seems that further fundamental research work on the properties of stabilized froths is urgently required. We have found another difficulty associated with the presence of very fine particles in phoscorite slurries; these particles can be easily floated but with very poor selectivity. We believe that it is possible that a purely physical mechanism such as entrainment is responsible for the collection of fine particles in the froth. We have also demonstrated that the presence of fine particles in the pulp adversely affects the rate of flotation of the larger particles. For esample, we have found an increase in recovery of + 100 um material from 2 percent to 17 percent in a standard laboratory batch flotation test on phoscorite when the - 37 um was removed from the feed. Such interactions between particles are not universally recognised as an important aspect of the complex flotation mechanism. They should appear in any quantitative formulation of the flotation rate process.

A General Meeting, followed by a Colloquium on &apos;Pilot plants and problems of scale-up&apos;, was held on 22nd March, 1978, in the Main Hall, Kelvin House, Johannesburg. The President, Mr P. A. von Wielligh, opened the meeting at 09h15. He thanked the authors and contributors for their co-operation, and announced the names of the undermentioned candidates, who had been admitted to membership of the Institute:

Jan 1, 1978

By A. K. Haines

Hydrometalurgy, particularly leaching, is a highly topical subject and always generates great interest in the mineral industry. This was evident right from the start of the organization of the SAIMM Colloquium on Leaching, which was held at Mintek on the 8th and 9th November, 1988. The response to the cal for papers was so good that the Colloquium was extended to two days, instead of the usual one day.

Jan 1, 1988