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By N. Andersen

During the preliminary design phases and feasibility study of a proposed vertical shaft, a vertical diamond drill-hole is normally drilled on the site. This paper outlines how rock mass characteristics, in-situ rock stress, rock strength, hydrological characteristics, and structural parameters can be determined using wireline logging of this borehole. In addition, a rock mass classification scheme is developed, based on published work, and in particular the Q-factor is adapted to assess the stability of raise-bored shafts. The ‘stick plot’ is introduced, which combines all geotechnical parameters applicable to the stability of a vertical shaft into a colourcoded format where cross-correlations can readily be made on a day-today basis during the shaft-sinking process.

Jan 1, 2015

By R. J. Dippenaar, J. C. van den Berg

This paper describes an investigation into the prereduction of hematite fines in a fluidized-bed reactor as a means by which they can be prepared for direct smelting in a transfered-arc plasma furnace. A single-stage fixed fluidized bed reactor using the in situ gasification of coal was employed. It is shown that the hematite is reduced to magnetite at 1000 QC, the degree of reduction being limited by the inherent shortcomings of a single-stage reactor in which both the gasification and the reduction reactions have to occur. Methods by which the degree of reduction can be increased are suggested. The gasification process was studied in some detail, and three different reducing agents are compared. The dominating effect of the Boudouard reaction on the behaviour of the fluidized bed is clearly illustrated. Non-coking coals of high reactivity are best suited to this process, a finding that would complement the trend away from coke as a metallurgical reducing agent.

Jan 1, 1989

By R. Katzenbach, M. Vogler, U. Adamietz

Abstract The design of deep excavation pits in rock mass is dependent as well on the mechanical properties of the rock mass as for example the orientation and the type of discontinuities like bedding planes, joints and fissures, the length of the discontinuities (persistence), the grade of weathering, the strength of rock mass, and the shear strength in the discontinuities, etc., and also on the respective boundary conditions specific for the project, and here especially the geometrical boundary conditions. The basic mutual influences and interactions between these boundary conditions are explained and exemplary presented at an excavation pit with a depth of 18,5 m, situated at a slope with a surface near groundwater table. Résumé La planification des excavations profondes dépend aussi bien de la qualité mécanique du roc, par exemple la location de la texture de la surface de moindre résistance, le degré de rupture, le degré de décomposition, la dureté du roc et la résistance limitée au cisaillement de la texture de la surface de moindre résistance que des conditions limitées spécifiques pour le projet, ici en particulier des conditions limitées géométriques. Les influences et interactions essentielles et mutuelles entre les conditions limitées isolées seront expliquées et présentées à l’aide d’une excavation 18,5 m de profondeur, située sur une pente où le niveau des eaux souterraines est près de la surface. Kurzfassung Die Planung tiefer Baugruben im Fels hängt sowohl von den mechanischen Eigenschaften des Fels wie z.B. der räumlichen Lage des Trennflächengefüges, dem Durchtrennungsgrad, dem Verwitterungsgrad, der Gesteinsfestigkeit sowie der Scherfestigkeit des Trennflächengefüges als auch von den projektspezifischen Randbedingungen hier insbesondere von den geometrischen Randbedingungen ab. Die wesentlichen gegenseitigen Einflüsse und Interaktionen zwischen den einzelnen Randbedingungen werden aufgezeigt und an einer 18,5 m tiefen Baugrube in Hanglage mit oberflächennahem Grundwasser vorgestellt.

Jan 1, 2003

By M. Matabane, T. Khati

"Accurate delineation of the contact between a kimberlite pipe and country rock at production level depths is a challenge due to limited geological data. Geological information is obtained from widely spaced diamond core boreholes which are drilled either from surface or from higher mining levels within the pipe. Kimberlite pipe/country rock contacts are notoriously irregular and variable, further reducing the confidence in contact positions defined by the drill-holes. At Finsch Diamond Mine (FDM), the opportunity arose to further improve the confidence in the contact positions relative to the planned slot (end) positions of each sublevel cave tunnel during the development stage of these tunnels. As a result, the accuracy of the 3D geological model has improved.The use of diamond drill core for this purpose is expensive due to site establishment requirements. The lengthy time taken during site establishment also delays the development of tunnels and support cycles, thereby extending the completion dates. FDM has reduced delays during development by adopting percussion drilling, in conjunction with gamma ray logging. The S36 drill rig is mounted on a moveable platform and does not require a costly and lengthy site establishment. The holes are generally drilled (0°/flat) on grade elevation, and these holes could also be drilled from the rim tunnels (developed in waste) into the kimberlite pipe. A single-boom production drill rig is normally used to drill holes about 20 m in length.On completion of the contact delineation drilling, gamma logging of the holes is conducted using the GeoVista geophysical sonde (or probe) to log the natural gamma signature of the dolomite/ kimberlite contact. The advantage of this tool is that the readings are continuous within centimetre intervals, and due to contrasting characteristics between kimberlite (rich in clay minerals) and dolomite, the contact position can be determined accurately. The better definition of contact positions also adds value to tunnel stopping distance in terms of developing the tunnel’s slot at the optimum distance from the contact (easier blasting of longhole rings, avoidance of contact overbreak and premature waste ingress, and other matters relating to extraction of ore from these tunnels). This method is highly successful and has reduced development costs (on-time completion), improved definition of the pipe’s contact position for geological modelling, improved blast design, and mitigated early waste ingress by maintaining the contact’s integrity."

Feb 1, 2019

By C. Musingwini, A. S. Nhleko, M. J. Mahase

"In an environment like the mining industry, which is characterized by different stakeholders with multiple objectives, multi-criteria decision analysis (MCDA) is a useful approach for optimal decision-making. The application of MCDA techniques in the mining industry has predominantly been in mine planning and related problems, although no comprehensive survey has previously been undertaken to establish the application trends. A survey of the use of MCDA techniques was therefore conducted using case studies from the literature. It was noted that often two or more methods are applied to the same problem in order to increase confidence in the solution derived. As the number of criteria and alternatives increases, some methods become inefficient. A combination of the analytic hierarchy process (AHP) method with other MCDA techniques was the most frequently used approach, indicating the efficiency of the AHP method, especially when evaluating problems with more criteria and fewer alternatives. A combination of fuzzy theory with AHP or other methods incorporates uncertainty. The findings from the survey will benefit users applying MCDA techniques to solve mine planning and related problems.IntroductionA country’s natural resources serve the needs of different groups of people within the country. These different groups are collectively known as stakeholders (Herath and Prato, 2006). Stakeholders have different needs, preferences, and expectations, which describe and define the criteria to be considered in decision-making. The options available as solutions describe the alternatives in decisionmaking theory. The criteria and alternatives have to be evaluated simultaneously when making decisions, and this is when multicriteria decision analysis (MCDA) techniques become invaluable (Herath and Prato, 2006) because the human mind is considered to be limited when solving complex and intangible problems with several criteria and alternatives (Saaty, 2007). MCDA techniques belong to a broad group of techniques called multi-criteria decision-making (MCDM) techniques. The other group of techniques within MCDM comprises the multi-objective decision-making (MODM) techniques. MODM techniques are generally used to solve problems with an infinite number of alternatives, while MCDA techniques are generally used to solve problems with a finite number of alternatives (Musingwini, 2010)."

Jan 1, 2016

By P. S. G. Leon

One of the most difficult areas in concession contracts in resource-rich developing countries is achieving an equitable balance between the interests of investors, host states and, increasingly, mine and near mine communities. While some countries in the Andean region have resorted to outright or creeping expropriation of natural resource contracts under the guise of resource nationalism, others such as the Democratic Republic of the Congo (DRC), have claimed international best practice in the review (and setting aside) of important concession contracts.

Jan 1, 2011

By S. Asokan, CVGK Murty, R. Upadhyay

The principal mineral sources of titanium are rutile, anatase (both TiO2), ilmenite (FeTiO3) and leucoxene (a weathered ilmenite of variable concentration of TiO2 but similar to pseudorutile Fe2Ti3O9). These titaniferous minerals along with their value added products like Synthetic Rutile and TiO2 slag constitute ?Titanium Feed-stocks? for TiO2 pigment, Ti metal and welding electrodes industries. Unlike in other industries, the demand driver for titanium minerals is not the metal, but the TiO2 pigment ?a specialty chemical. Of the total titanium feedstocks usage in the world, about 93 % is consumed in TiO2pigment, with only about 3 % is utilized in metal production. India?s heavy mineral sands resources are among the largest and also one of the richest grades in the world. The country?s Ilmenite resource base stands at 348 million tonnes of Ilmenite (18% of the total world reserves)along with 18 million tonnes of Rutile, 21 million tonnes of Zircon, 8 million tonnes of Monazite, 107 million tonnes of Garnet and 107 million tonnes of Sillimanite. Though the grades (Total Heavy Minerals (THM) present in the sand) of Indian deposits are relatively rich, the contained ilmenite are termed as sulphate grade with 50-53% TiO2 (Chavra deposit of Kerala and Ratangiri deposit of Maharastra are an exception with 60% of TiO2). While superior grades with adequate reserves lead to a cost advantage in ilmenite production, the lower TiO2 percentage allows the direct use of Indian ilmenite only in the production of TiO2 pigment through sulphate route and for the chloride pigment production; such ilmenite is required to be upgraded either to Synthetic Rutile (SR) or TiO2 slag. However, in order to produce TiO2 slag in India at a globally competitive cost, the issues like high power tariff, inadequate infrastructure, expensive logistics, less enthusiastic market preferences, supply-demand gaps are to be adequately addressed. This paper deals with some of the above issues and possible solutions thereto in production of TiO2 slag from Indian ilmenite.

Jan 1, 2007

By K. L. Morton

This technical note reviews the status quo of mud rushes and flooding, which have claimed them lives of many people in underground mines in South Africa and globally. Given that the causes are well known and can be managed, the note calls for the development of a national and global standard on flood and mud rush prevention in order to save lives.

Mar 30, 2026

Dr Hanekom I would like to thank Dr Hanekom for a very valuable -contribution; it summarizes careful statistical test work which was an omission from the original paper. Minor manipulative details, the application of a finer measuring graticule and the reproducibility of the procedure are described and the author accepts with gratitude this ready-made answer to the first and second points of Dr Finkelstein's comments. Mr Williamson The author is indebted to Mr Williamson for historical background to early experiments with captive bubbles and it is conceded that the original paper described another avenue to an old flotation goal. The field of application is restricted and we would not work with all the material illustrated by Klassen and Mokrousov.1 The particles in Figure 74(a) are far too large and irregular; material in Figures (b) through (d) could be employed for semi-quantitative work in spite of obvious disadvantages; Figure 74(f) represents ideal conditions. We have studied this mineral-collector association in the size range given and results were satisfactory. For larger pick-up values than that illustrated an induction period of more than 0.01 second was required (actually 90 seconds for a maximum bubble load at pH 7.2 with 25 g/t oleic acid!). REFERENCE 1. KLASSEN,V. I. ANDMOKROUSOV,V. A., 'An Introduction to the Theory of Flotation,' Butterworths, London, 128, Figure 74, (1963). Mr Slander and Mr Kooij Researchworkers in a practical mineral development laboratory may well ask what contribution yet another micro-technique can make to the problems of large scale ore flotation. The obvious answer is that there can be little or no direct translation of results obtained with a static system comprising air bubble and pure mineral concentrate to a dynamic multicomponent operation. However, for the rapid study of one mineral or variable some small scale apparatus must be employed. In the Anglo American Research Laboratory, the described semi-quantitative pick-up technique has filled a niche and requirement. For ourselves, the system represents a definite advance over the contact angle approach in that samples of minerals conventionally prepared in laboratory or plant may be used, which is a desired step nearer flotation. We bracket this pick-up apparatus with the Hallimond tube1,2 and evidence is now to hand that pick-up curves may often be related to flotation recoveries obtained from this tube with pure mineral samples and from the Fuerstenau glass cell3 with the total ore at low pulp densities. We would not wish to push the comparison of pick-up results past this stage for there is much published development work relating the Hallimond and Fuerstenau equipment to flotation on a larger scale. At the Anglo American Research Laboratory for example, a test in a Fuerstenau cell is often the final integrated step in laboratory investigations into the TORCO process. As these contributors rightly point out, pick-up tests can offer nothing towards the determination of optimum reagent dosage for large scale flotation. This is the function of bench-type flotation cells with up to 2 kg batches of ore. At this level, collector performance and consumption and physical properties of the froth are more comparable with large scale operation. The pick-up cell, Hallimond tube and Fuerstenau cell are essentially equipment from which the effect of control parameters can be derived. For the assessment of collector behaviour, the pick-up approach is rapid and economic and we offer here a pick-up contribution towards the flotation of Witwatersrand uraninite with anionic sulphonate collectors (Fig. 1). A large crystal of the local mineral was not available for contact angle measurements and owing to its low concentration in Witwatersrand gold-bearing reef, a small sample was laboriously obtained by tabling, superpanning and heavy medium separation. The pick-up values of a sized portion were determined by the described procedure in the presence of nine petroleum sulphonate collectors. Responses fell into two categories; "normal" curves (dotted lines) and a group of collectors which were very active at low concentration but showed depression of pick-up at the higher levels. The similarity in form of curves for collectors 1 and 3 was apparent and reagent 3 was found to be simply a diluent of 1. Otherwise, collector performance could not be related to either sulphonate content or molecular weight of the active constituent. However, reagent 2 gave the best performance in a bench flotation cell with regard to overall uranium distribution but, as predicted, reagent 1 yielded the best grades. Later pick-up tests with a similar class of collector have suggested that depression at high reagent concentrations may be due to "bubble armouring" whereby an hydrophilic air-liquid interface is produced which rejects mineral attachment. Prolonged investigations with the same mineral sample have not been required except for the study of uraninite which lasted for more than a year. Variation in pick-up response could not be detected over this period during which the mineral was stored under distilled water. The only visible change was a small efflorescence of lead sulphate, readily removed and decanted off after stirring with a glass rod in the sample tube. In its main role as a "trouble-shooter", the pick-up method is applied to fresh mineral samples taken from a flotation plant and soon discarded or from a current run in the ore-dressing laboratory. After production of a concentrate, a solvent cycle, as described in the original paper, may be required to remove heavy organics or other contaminants as when a sample has been taken from circuits containing detergents, grease or return flotation solutions. This solvent cycle will not remove chemisorbed collectors and we prefer not to work with minerals from flotation tailings or with pick-up cell discards. It has been thought preferable to answer the contributors' other apprehensions with regard to sample preparation, activators, depressants and the physical properties of reagents in another paper on offer to this Journal. It covers pick-up investigations relating to the flotation of nickeliferous pyrrhotite at two Anglo American Corporation Mines, which approach seemed the correct orientation for a reply to another contributor to this paper.

By Y. Sagawa

The Hishikari Gold Mine consists of epithermal vein type Au-Ag deposits, each of which has several veins standing almost in parallel each other and inclining 70 to 90 degrees. The veins have 1 to 3m widths and are extracted mainly by drifting and bench stoping with waste rock fill. The Hishikari has tried to optimise the stope dimensions by the Potvin’s Stability Graph Method, which is well known as an easy and effective tool to design appropriate stope dimensions. The method, however, has some limitations of application due to its simplicity so that it has been used only in limited cases. The authors developed a new numerical approach, based on 2D FEM solu-tion for the stability analysis. As a result, it has been possible to solve the Hishikari situation. The authors also introduce the concept of the design approach and compare the results obtained by us-ing the approach with the database of the Stability Graph Method for validation. La mine d’or Hishikari se compose de gisements d'Au-AG avec des types de filons épither-maux dont chacun a plusieurs veines qui sont presque parallèle les unes par rapport aux autres et sont inclinées de 70 à 90 degrés. Les veines ont de 1 à 3 m de largueur et elles sont principalement extraites en creusant et en abattant des bancs de roches encaissantes. Hishikari a essayé d’optimiser les dimensions du chantier d’exploitation avec la méthode de graphique de stabilité de Potvin, bien connue comme un outil facile et efficace pour avoir des dimensions de chantier d’exploitation appropriées. Toutefois, la méthode a certaines limites d’application dues à sa sim-plicité de sorte qu’elle a été utilisée seulement dans des cas limités. Les auteurs ont développé une nouvelle approche numérique, basée sur la solution 2D FEM, pour l’analyse de stabilité. Par conséquent, il a été possible d'expliquer la situation de Hishikari. Les auteurs présentent le concept de l'approche et ils comparent les résultats donnés en l'employant avec la base de données de la méthode de graphique de stabilité pour la validation. Die Hishikari-Goldmine besteht aus Au-Ag-Ablagerungen mit epithermalen Adern, die jeweils mehrere Adern aufweisen, die mit einer Neigung zwischen 70 und 90 Grad nahezu parallel zueinander verlaufen. Die Adern haben eine Breite von 1 bis 3 Metern und werden überwiegend durch Vortrieb von Stollen und Bergversatz abgebaut. Hishikari hat versucht, die Abmessungen des Abbauorts durch das Potvins Stabilitätskurvenverfahren zu optimieren, das als einfaches und effizientes Mittel für die Erzielung angemessener Abmessungen des Abbauorts bekannt ist. Das Verfahren weist jedoch aufgrund seiner Einfachheit einige Anwendungsbeschränkungen auf, so dass es nur in einigen Fällen benutzt worden ist. Vor diesem Hintergrund haben die Autoren einen neuen numerischen Ansatz entwickelt, der hinsichtlich der Stabilitätsanalyse auf der 2D-FEM-Lösung beruht. Dadurch wurde es möglich, das Problem von Hishikari zu lösen. In diesem Dokument stellen die Autoren die Grundlagen des Ansatzes vor und vergleichen die erzielten Resultate durch Benutzung der Datenbank des Stabilitätskurvenverfahrens zur Validierung.

Jan 1, 2003

By A. Ochieng

Solids suspension influences the quality of mixing and energy requirement in a solid-liquid system, both of which determine the efficiency of industrial processes such as nickel precipitation. Nickel solids concentration distribution in a stirred tank was investigated using computational fluid dynamics (CFD) and experimental methods. The concentration distribution of the nickel solids was compared with that of sand and glass. The laser Doppler velocimetry (LDV) method was used to measure the velocity field for the liquid-only system and an optical technique was employed to determine the axial solids concentration distribution. Regions of in homogeneity in the tank were identified. It was found that, for a given solids loading, the solids concentration distribution depended on both particle size and particle size distribution. High solids loadings were investigated and a difference in the concentration distribution pattern was obtained with nickel, flint glass and sand particles. The CFD simulation results highlighted problems that could be associated with some conventional experimental methods of determining solids concentration distribution in a stirred tank. Keywords: nickel; CFD; simulation; mixing; solids suspension.

Jan 1, 2010

A shortage of technical skills in all areas of the mining industry is being experienced worldwide. This has significant consequences for design, operations, productivity and safety and, therefore, needs to be seriously addressed, both in the short and longer term. It is therefore both a strategic and a tactical issue. Importantly, however, since mining is a long-term business, the shortage of skills must be dealt with from a strategic point of view to ensure the supply of the necessary skills for future mining. This paper will deal specifically with the provision of high level technical skills, that is, university graduates in mining engineering, and the outlook for the institutions that are currently producing such graduates.

Jan 1, 2008

By Hornsey R., P. van der Merwe

The Two Rivers Project is a feasibility study to evaluate the UG2 chromitite layer on Dwarsrivier372KT, Mpumalanga Province, South Africa. The project, which commenced in August 2001, is expected to receive go-ahead early in 2003. The project sponsors (AVMIN 55%, Implats 45%)required project delivery within a short timeframe requiring optimization of the geological and geostatistical programme. Geological modelling was carried out in parallel to the drilling programme, and significant up-front geostatistical preparatory work was done. The time saving was approximately 4 – 6 months on project delivery. The cost saving to the project (excluding NPV considerations) is estimated at R20–30 million. Keywords; Two Rivers Platinum, Dwarsrivier 372KT, UG2 chromitite layer.

Jan 1, 2003

By A. J. Jager, D. P. Roberts

The stoping of platinum orebodies in the Bushveld has in the past been accomplished by using in-stope pillars to ensure stability of the workings. In the future, the use of continuous mining machines to mine the platinum orebodies is probable. It is not envisaged that such machines will cut in-stope pillars. This paper describes the alternative support systems that will replace the pillars and the stope geometries that will be required for this type of mining. The paper describes how the support resistance provided by in-stope pillars in current mining was determined and then evaluates support systems that can substitute for these pillars. Conceptual layouts based on existing breast mining layouts are presented. The layouts consist of panels of up to 100 m face length between rigid strike stabilizing pillars. A number of design charts are then proposed that allow for the design of appropriate support systems.

Jan 1, 2004

By P. R. Jochens

South Africa depends to a considerable measure on the exploitation of mineral reserves and the processing of these minerals. One of the prime requirements for the mining and metallurgical industry to continue in this vital role is access to sufficient energy in a suitable form. Hence, four important questions are discussed: Does South Africa have sufficient energy for the processing of its minerals and metals! Will the cost of energy allow South Africa's minerals and metals to compete on world markets? Can mining and metallurgical processes become more energy-efficiency Can mining and metallurgical processes be adapted to the various forms of energy that may have to be utilized in the future? It is concluded that about 26 per cent of the extractable reserves of coal would be required in the mining and processing of 50 per cent of South Africa's reserves of gold, platinum-group metals, copper, iron, ferrochromium, and ferromanganese. However, the reducing agents required exceed the extractable reserves but not the total estimates of mineable in situ resources of metallurgical coal and anthracitic coal. Even under the severe constraint that coal could become virtually the sole source of energy, only about three-quarters of the extractable reserves of coal would have been consumed by the year 2025. The reasonably assured resources of uranium metal that can be recovered at less than $80 per kilogram are considerable and, at current conversion efficiencies to electricity, are equivalent to one-fifth of the extractable reserves of coal. The rate of exportation of energy in the form of U30. was about five times that of coal in 1978. It is pointed out that the price of energy has a direct effect on the competitiveness of metals, and, although constituting only a small proportion of the selling price of precious metals, it represents a significant proportion of the selling price of copper and metals recovered from oxide ores. The scope for the conservation of energy appears to be greater in the extraction and conversion of primary energy-carriers to electrical energy than in other mining or metallurgical fields. Provided that alternative sources of energy can be converted to electrical energy, they can be applied to the mining and metallurgical industries. The decision as to which type of energy should be used in specific mining and metallurgical situations is complicated, and is likely to be subjected to closer scrutiny than in the past. The hope is expressed that the tentative conclusions reached will give rise to more-quantitative analyses from those qualified to undertake such an exercise. Inevitably, the questions asked at the outset give rise to more questions, and the more interesting ones are posed in the context of each conclusion.

Jan 1, 1980

By A. M. Hefny, F. P. Siew, P. G. Ranjith

Fracture initiation and propagation, being one of the most intensive subjects in rock mechanics; contribute significantly to the deformation of jointed rock masses. An experimental study on fractured granitic rock was carried out to determine the peak strength of specimen under triaxial loading conditions. The commonly observed failure mode for rock specimens in this study was found to be the shearing of the plane of weakness. The minimum peak strength of fractured rock is observed when the joint orientation is approximately 70° to the horizontal axis. Based on experimental test data, an empirical expression was developed to accommodate the effect of joint orientation and joint trace length in estimating the peak strength of fractured rocks. The predicted values of peak strength using the proposed equation well agree with the experimental results carried out on singly fractured specimens under triaxial loading conditions. The threshold stress values of crack initiation and propagation depend on joint geometrical parameters, their degree of interconnectivity, as well as surrounding stresses on the fracture plane.

Jan 1, 2003

By More-O&apos, R. C. Ferrall, J. W. Wilson

Discussion: M. D. G. Salamon* (Fellow): The theoretical principles behind the techniques of calculation used by the authors was formulated and published during the period 1962- 65. These principles, together with their practical implications, were summarized in a review paper in 1966.1 Now, in February, 1970, we have this long and comprehensive paper discussing the practical application of the theory. I think this is good progress. The large number of practical examples treated in the paper clearly demonstrates not only the skill and knowledge of the authors, but also the versatility of the techniques of analysis. I feel that the techniques described by the authors, and their improved versions, will revolutionize mine planning. I suggest that no shaft system or stoping layout should be planned in future without an adequate rock mechanics analysis. It seems that it will be progressively more difficult to white wash mistakes in rock mechanics and attribute their costly outcome to normal mining risk. We should not, however, become complacent. There is still a lot to learn and do in the future. I do not want to discuss tonight the question of further research but instead, I will raise two other problems which may hinder the future progress of practical rock mechanics. To carry out work similar to that described in the paper, mine managements require men who are capable of doing the work and who have the most efficient techniques of analysis at their disposal. As a research engineer I have been associated with the development of techniques of analysis. I can assure you that these have undergone an amazingly rapid development during the last six years. We have seen the change from the use of the electrolytic tank analogue and manual integration to the method described by the authors, that is, to the employment of the automatic network analogue and computer integration. But this is not the end. In a recent paper Prof R. P. Plewman2 and his co-authors described a completely digital technique. Here the process of calculating on- and off-reef quantities is integrated into one computer run. But these sophisticated methods of computation are useless without men to make intelligent use of them. During the last few years a new breed of mining engineers has appeared on the scene - the Rock Mechanics Engineer. I feel it is timely to call the attention of the senior members of this Institute to the fact that this new breed will succeed in carrying out its duties in a valuable manner only if we manage to attract to its ranks young engineers of the right quality. The requirements are high. The top rock mechanics engineers have to be good mine planners and they have to understand the essence of sophisticated theories. Without these attributes they will not succeed in the long term and the industry will not gain the benefit of available knowledge. To attract this quality of engineers the industry will have to ensure that rock mechanics is accepted by all concerned as a career which could lead to the top echelons of our profession. Finally, I would like to make one or two remarks concerning Part I of the paper. In Sections 3.1 and 3.2 the authors describe methods of determining the normal stress on the reef plane and the convergence distribution in the excavations. The description as given applies only to a horizontal reef. When the reef is inclined the situation is more complex since a shear component of stress on the intact reef and a ride component in the excavations must then also be determined. These can be calculated by methods similar to that described in the paper.3 The practical application of the method of calculating the released energy (Section 3.4) is in the comparison of various stoping layouts to establish an order of preference in terms of decreasing rates of energy release. To carry out these comparisons effectively, the calculations must be carried out in a manner by which the values of energy per unit area obtained are comparable. I would like to note in this respect that the method of calculation given in the paper is valid without reservation only if there is no elastic contact between the hanging and foot walls in the mined-out area. If there is contact then the calculated energy values are correct only if the mined out area is increased in small steps during the analysis. My second point in this regard is concerned with the practical method of calculation as described in Sections 3.4 (ii)-(iii). Firstly, I would like to suggest that the energy calculation should be carried out, as far as possible, by using always the same scale on the analogue. Secondly, the reading of current on a pin should be followed by the removal of that pin to obtain the corresponding reading of potential. The product of these two values will be proportional to the energy released during the mining of the small area corresponding to that pin. The method suggested by the authors tends to mask possible danger points in the layouts, since they obtain an average value for a time period, say, for six months.

By G. L. Smith

This paper introduces and defines the concept of a linkage between long-term planning parameters (?global assumptions?) and possible future world views as conceived in scenario planning. Scenario planning, the tool that is used to arrive at plausible future worldviews, is described in the context of the platinum industry. The interpretation of the scenarios into long-term planning parameters is made, followed by an illustration of how the long-term planning process has been adapted at Anglo Platinum to include the notion of uncertainty.

Jan 1, 2008

By G. C. Lindeque

"A downward trend in tyre life on haul trucks at an open pit iron ore mine prompted an investigation into the current tyre management strategy in order to improve tyre life. ‘Best practice’ was compared to the current tyre management strategy at the mine to determine aspects that could be improved. Aspects that showed significant room for improvement included employee tyre awareness, tyre pressure maintenance, and road conditions. For each of these aspects a plausible alternative was suggested.A new tyre management strategy was formulated as a step-by-step guide on how to implement the proposed changes. Each step introduces new initiatives but at the same time builds on the previous step’s success. A sensitivity analyses was performed to determine the impact of the proposed changes, with the variables being the reduction in the percentage of premature tyre failure, and the reduction of the site wear rate.The ‘most likely’ scenario showed an increase in tyre life of 41% for the Komatsu 730 trucks and 105% for the CAT 777 trucks if the entire tyre management model is implemented. This would result in a possible annual saving related to tyre usage of R9.2 million. OverviewThe study focused on the tyre life on the main haul fleet at an open pit iron ore mine. The haul fleet consisted of 17 Komatsu 730 (180 t) haul trucks and 12 CAT 777 (90 t) haul trucks.The haulage operation consists entirely of tyre-based haul trucks. Five per cent of the total mining budget is related to tyre costs. Since 2012 there has been a steady decrease in tyre life as measured in operating hours (Figure 1) due to the high percentage of tyres that fail prematurely. In 2014 a total of 41% of all tyres failed prematurely and 61% were classified as ‘worn out’."

Jan 1, 2016

By M. K. G. Vermaak

Spiral concentration forms a crucial part of gravity separation circuits. The application of process models in the operation of gravity separation circuits offers significant benefits in recoveries and grade control. Significant strides in the modelling of spiral concentrators have been made during the last couple of decades. Despite these advances in modelling techniques and models, application in industrial processes is not fully optimized. The adjustment of splitter positions to accommodate changes in the feed parameters is impractical and difficult, resulting in losses. In this work an optical method is employed to detect the concentrate band position as a function of operating variables. As expected, the results indicate a strong influence of the solids content and total heavy mineral (THM) feed grade on the predicted gangue-mineral interface. Viscosity modifiers indicated the band position to be sensitive to viscosity and therefore the slimes content. Validation on an industrial sample indicated the possibility of using this model in feed-forward control application.

Jan 1, 2008