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By J. Kortnik

Mining has always been, and still is, an activity which is vital for the development of mankind. This is because a predominant proportion of industrial products and energy is produced from raw materials that are excavated using mining methods. The need for environmental protection and the close proximity of urban settlements and mines in Slovenia have necessitated ever stricter requirements for the excavation of raw materials and introduction of more environmentally-friendly mining methods. This is achieved by implementing a closed eco-technological cycle of raw material excavation within the scope of mining methods. In surface mines, this means prompt reclamation of the degraded surfaces, and in underground mines the use of methods with minimized impact on the surface and on underground water. In underground mines, mining methods with backfilling installed according to the multi-barrier disposal system are the only ones that allow the possibility of implementing a closed eco-technological cycle of excavating raw materials. Various industrial waste materials are also included in the backfilling material in order to improve its strength properties. Backfilling consisting of several different waste materials is produced in the form of a composite. Before backfilling consisting of waste material composites can even begin to be used in mines, an environmental impact assessment of its suitability for mines needs to be made. In order to make such an assessment, a pressure leaching cell and an open diffusion cell were made, i.e. a pressure leachate test and a sorption test were developed. The two cells and the developed tests will enable the study of the influence of backfilling on the environment if installed in an underground mine. The pressure leachate test performed with the use of the pressure leaching cell simulates leaching from a consolidated backfill during the seeping of underground water into the backfilling material under pressure. The sorption test performed in an open diffusion cell simulates the flow of eluate leached from the backfilling material through the rock/geotechnical barrier. The two tests were developed in such a manner that they are performed either in succession immediately after one another, but they can also be performed separately at a certain time interval. The paper presents an assessment of the suitability of backfilling made of different waste material composites. Keywords: environmental impact assessment, multi-barrier disposal system, waste material composites, pressure leachate test, sorption test, pressure leaching cell, open diffusion cell, eco-technological cycle, mining methods with backfilling.

Jan 1, 2003

By J. Wills

The use of drilling rigs to increasingly mechanize stope drilling has been acknowledged as important for the S.A. mining industry in it ís efforts to tackle the challenges of low productivity, AIDS and an ageing driller workforce. The spin-offs of increased safety, cost-effectiveness and better mining performance provide advantages to all?from the driller to the shareholder. However, implementation or product rollout has varied from outright success to dismal failure. A study of various cases by Novatek and Anglogold TDS showed clear reasons for these differences. Whereas the successes were characterized by clear objectives, accountability, user buy-in, a commitment to make the technology work and no-nonsense reporting, these were typically absent in failed implementations. The attitude at unsuccessful mines was to ?trial? the equipment?even subconsciously, a ?trial? is a test where failure becomes an acceptable outcome. During the study it became clear that mining management are not adequately trained to manage the often-tricky rollout process. They are often also not empowered or clearly mandated to deal with the organizational, human and technical issues that must dovetail to create a success. The output of the study was to develop a basic rollout process that could be used to formally manage the implementation of new drilling technology. Creating an alignment between the mine, workforce and supplier promotes a win-win situation that will not otherwise spontaneously arise.

Jan 1, 2002

By R. Poisel, C. Scavia, Preh A., W. Roth, M. Pirulli

ABSTRACT: Two methodologies for the three-dimensional simulation of the run out of rock avalanches are presented: one based on a granular approach and the other based on a continuum equivalent approach. An evaluation of the possible run out area of the Oselitzenbach landslide (Carnic Alps, Austria) has been carried out applying the two methodologies. The combined use of these two approaches makes it possible a comprehensive assessment of the run out area. RESUME: Deux méthodologies pour la simulation tridimensionnelle de la propagation des avalanches rocheuses sont présentées: la première est basée sur un modèle de type granulaire et la deuxième sur un modèle du continu équivalent. Une évaluation de la propagation éventuelle de l’avalanche de Oselitzenbach (Alpes Carniche, Autriche) a été réalisée en utilisant les deux méthodologies. Les résultats obtenus montrent que la combinaison des deux approches rend possible une analyse plus complète de la phase de propagation. KURZFASSUNG: Für die dreidimensionale Modellierung eines Felsmassensturzes (sturzstrom) werden zwei Methoden, eine granulare und eine kontinuumsmechanische Methode, vorgestellt. Beide Methoden wurden zur Abschätzung der Ausbreitung (runout) der Massenbewegung Oselitzenbach (Karnische Alpen, Österreich) herangezogen. Die kombinierte Anwendung beider Methoden ermöglichte eine umfassende Beurteilung des möglichen Ausbreitungsbereichs.

Jan 1, 2003

By J. R. G. Williamson

The Mining Division of Iscor, an integrated minerals and metals company, has recently completed a feasibility study for the establishment of a new heavy minerals mine and smelter in the Richards Bay area of KwaZulu-Natal, South Africa. The mine will produce a heavy mineral concentrate containing ilmenite, zircon, rutile and leucoxene which will be beneficiated at a central processing complex close to Empangeni. The ilmenite will be smelted to produce high-grade titania slag and pig iron for export. The ancient dune containing the mineral deposit is located some distance inland of the coastal dunes, and the sands contain a high proportion of extremely fine material which must be extracted ahead of the gravity separation process. Of the 56 million tonnes to be mined annually over the ten-year life of the mine approximately 23,4%?a total of 13,1 million tonnes of minus 45 micron residue fines?will have to be disposed of in a containment facility. The origin of the fine material is from the weathering of amphibole minerals. The remaining coarse sands will be returned to the mined area to re-create the original dune configuration. A number of options for the disposal of the residue fines have been examined and specialised laboratory tests undertaken to determine the deposition behaviour of the fines. The final solution involved the use of a ?sub-aerial? technique for deposition of fine material in a disposal facility. The resulting arrangement will adequately meet technical and environmental criteria. Upon closure after 10 years of operation, the surface of the disposal facility will be returned to sugar cane cultivation The paper refers to the various options considered and describes the problems associated with competent disposal of the residue fines and the development details of the final disposal facility.

Jan 1, 1997

By L. A. Cramer

Platinum metal was first utilized by the Inca civilization in South America where it was found in alluvial deposits along the San Juan and Atrato rivers that lie between the Andean mountain range in Colombia and the Pacific Ocean. The nuggets found in these rivers average between 77% and 86%platinum, 7%?8% iron and small amounts of copper, gold and other Platinum Group Metals (PGMs). The Incas often alloyed platinum with copper to increase its workability and fabricated various ornaments from it. With the arrival of the Spanish in the early 1700s, platinum was labelled a nuisance for the Spanish in their quest for alluvial gold because it concentrated with the gold and had to be removed by either extended amalgamation or by hand sorting. It was at this time that platinum received the root of its current name for the Spanish referred to it as platina del Pinto, platina being the somewhat derogatory diminutive of plata (silver). Pinto was a river in Colombia where the Spanish first distinguished it. The English eventually adopted ?platina? and it is only in recent times that the Latin name platinum has been accepted as the metal?s common English name. From its first introduction to European scientists in 1740, platinum was unappreciated because of its brittle nature and difficulty of working and as a result had little commercial value. Its ability to be dissolved in gold did lead to some counterfeiting of gold bullion and even to the debasement of gold currency.

Jan 1, 2000

By F. -Z. Ji, A. G. Irons, M. Barati, K. Coley

In order to optimize carbon injection into oxidizing slags, in terms of carbon and iron yield, we must have a detailed understanding of the kinetics of the reactions involved. The current study measures the rate of carbon oxidation during injection at steelmaking temperature by means of injections of coal into EAF slags. The compositions of analysed outgoing gas were used to calculate gas flow rate and hence the gasification rate of carbon. A fundamental equation to predict residual carbon in slags was developed. The predictions using this model at moderate injection rates are in good agreement with experimental data. The reaction rate of EAF slag with carbon is under mixed control by the slag gas and the gas carbon reactions. The relative importance of each step depends mainly on the v-ratio of slags. A relative reactivity parameter ? has been defined and it is found to be important to carbon-slag reaction kinetics and slag foaming. Key words: carbon-slag reaction, kinetics, slag foaming.

Jan 1, 2004

By R. E. Wilmot

The problem of corrosion of reinforcement in concrete structures is internationally recognized and represents the single biggest expenditure in the preservation of steel reinforced concrete structures. While the corrosion protection of the reinforcing steel is no substitute for design and the placement of good quality concrete, it does provide for a viable and economical extension of structural integrity and service life in the day-to-day and practical environment in which concrete structures are constructed. The mechanism of corrosion attack and consequent deterioration of reinforced concrete structures is well understood and documented. New structures can be designed and built to withstand corrosion. However, due to short-term economic restraints, many new reinforced concrete structures continue to suffer as a result of corrosion damage, which results from the corrosion of the steel reinforcing. This paper explores the economic and practical benefits offered by the application of hot dip galvanizing as a viable means of protecting reinforcing steel concrete structures. The paper also discusses technical misconceptions that exist about application and performance provided by the use of hot dip galvanized reinforcing steel.

Jan 1, 2007

By M. Stephton, L. A. Cornish, R. Süss, N. B. Maledi, J. H. Potgieter, L. Chown

In today?s gas turbines, the emphasis is on saving energy and reducing the amounts of pollutants emitted. This can be attained only by designing alloys with higher melting points and the capability to retain mechanical integrity at increased temperatures. The development of nickel-based superalloys (NBSAs) has reached its limit, as the alloys are operating at critical temperatures close to their melting points. Further increase in the operating temperatures will result in dissolution of the strengthening phases and even melting. Recent trends are towards developing materials that are adequately strong but have much higher melting points and corrosion resistance than the NBSAs that are currently in use. Studies conducted on Pt alloys with similar structure to the NBSAs have presented mechanical results that are promising. This has led to further development of these alloys as a replacement for NBSAs. The operations of gas turbines are usually limited by hot corrosion, which could lead to catastrophic failures. Increased temperatures have an adverse effect on the corrosion of the alloys by making the environment more aggressive. Hence, during a material selection process, the corrosion behaviour of any selected material for a highly specialized application such as gas turbines needs to be assessed. Generally, coatings are applied to combat degradation by salt deposits, hence coated and uncoated NBSAs were selected as benchmark alloys. Corrosion mechanisms and kinetic studies of the Pt alloys were conducted. High temperature hot corrosion (HTHC) tests were conducted at 950°C for540 hours. Analysis of the corrosion products was conducted using a Raman spectrometer, scanning electron microscope with energy dispersive X-ray spectroscopy (EDS) optical microscopy, and X-ray diffraction study of the phases that formed. Both coated and uncoated NBSAs failed under the test conditions. Samples were covered with a porous non-adherent greenish layer which is characteristic of NiO. Pt alloys displayed superior corrosion properties compared with the benchmark alloys.

Jan 1, 2006

By P. M. Mashaba, S. O. Bada

An advanced high internal phase water-in-oil (HIP W/O) emulsion binder and flotation with an internal dispersed water phase of 95 vol% was used to beneficiate South African weathered coal from discard dumps. Five coal samples with calorific values ranged from 8.48 MJ/kg to 20.94 MJ/kg were utilized. The performance of the HIP W/O emulsion was compared to that of kerosene in terms of the physicochemical properties of the clean coal products. A 5 kg/t emulsion binder addition produced a high-quality coal concentrate with 11.76% ash and a heating value of 28.61 MJ/kg from a sample containing 29.34% ash. Using the same dose and coal sample, kerosene resulted in a clean coal concentrate with 17.47% ash and a calorific value of 26.42 MJ/kg. Further samples were also beneficiated using 5 kg/t of W/O emulsion. The highest increase in calorific value was achieved from the beneficiation of a sample containing 48.71% ash, from 8.48 MJ/kg to 18.16 MJ/kg. It has been established that emulsion binders can upgrade coal samples dumped many decades ago.

Aug 2, 2022

By A. Golder, S. P. Dandapat, I. Roy

One of the consequences of large-scale opencast mining for coal is the disposal of an enormous volume of overburden material that needs to be removed. With the enlargement of opencast mines and associated high stripping ratios, the volume of overburden extraction is rising. Overburden from the initial mine cut has to be necessarily dumped outside the mining area until enough void is created inside the mine for back-filling operations. The quantity of overburden placed outside the mining area normally varies from 10% to 30% of the total overburden removal. The stability of these dumps placed outside the mining area is of prime importance from the point of view of smooth operations of the mines and safety of persons in and around the dumping area. Owing to increasing pressure on land use, mine operators do not have any other option other than to place the waste dumps very near to the surface edge of the pit, which may lead to failure of the topsoil benches of the working pit. This study identified major controlling parameters that influence a stable but economic combination of height and slope angle of external dumps located close to the edge of an open coal pit.

Nov 27, 2024

Escalating energy and electricity costs have become one of the largest drivers of expenditure in mining operations. Over the last eight years, energy costs have tripled when expressed as a percentage of total expenses in South African mines. In an effort to manage and reduce electricity costs, energy management strategies can be developed, inefficient operating units replaced, and the operation of energy-consuming components of ventilation systems optimized. Power consumption on mines is controlled mainly by three strategies, namely load clipping, by which energy use is reduced for certain parts of the day; load shifting, by which energy use is shifted to other parts of the day; and energy efficiency, by which energy use is reduced permanently. In this paper several projects that were implemented using the first two strategies of load clipping and load shifting are investigated. The actual and potential savings that can be achieved by implementing such energysaving interventions are presented. To reduce the operating costs of ventilating and cooling underground mines permanently, system optimization studies must be completed. Methods that can be used to reduce energy usage by optimizing cooling and ventilation systems are described, and network simulation models that accurately reflect the current and planned ventilation conditions are discussed. These models are then used to examine various options for improving the overall ventilation and cooling strategy. Different optimization scenarios can be simulated, and this assists the design engineer in obtaining the most energy-efficient system that will satisfy design workplace conditions. The final outcome is a reduction in operating costs, which can result in better operating margins and an extension of the life of mine.

Jan 1, 2014

By C. Lottering

Copper sulphate solutions are produced during the pressure leaching of first-stage leach residue in a typical base metal refinery process. Apart from impurities such as selenium and tellurium, this leach solution also contains other precious metals (Rh, Ru, and it) due to dissolution in the pressure leaching stage. Selenium and tellurium are removed from the leach solution by precipitation with sulphurous acid, since these elements negatively affect electrowinning. This study investigates the feasibility of modifying the Se/Te precipitation process step to achieve significant rhodium precipitation in addition to selenium and tellurium precipitation, in order to reduce the rhodium inventory of the process. Thiourea, S02, formaldehyde, formic acid, and sodium thiosulphate were evaluated to determine which of these reagents would achieve the highest Rh precipitation. Based on these screening tests, SO2 and thiourea were selected to be used as precipitation reagents in optimization tests. During these optimization tests, the operating temperature (80 to 150°C) and the amount of reagent added (80 to 140 per cent excess) were varied to evaluate the effects that these operating conditions have on the precipitation behaviour of Rh as well as Se, Te, Cu, and Ni, and to propose appropriate operating conditions. It was found that thiourea resulted in good Rh and Se precipitation (typically in excess of 90 per cent), but precipitated noticeably more copper and nickel from the solution than when SO2 was used. In addition, thiourea had poor Te removal characteristics. SO2 achieved a maximum of 70 per cent precipitation of the rhodium. Faster kinetics and a larger extent of Te precipitation were, however, observed when using SO2 compared to thiourea.

Jan 1, 2012

By P. J. D. Lloyd

The widely employed IPCC models for the release of methane from underground coalmining operations are reviewed, and put in the context of the South African National Communication on greenhouse gas emissions. Then attempts to calibrate the model via measurements on South African coalmines are described. These attempts failed because it rapidly became clear that the IPCC model, involving methane adsorbed by the coal and released (a) by the depressurization of the strata caused by mining, (b) from the broken coal and (c) from coal remaining in roof, floor, pillars or goaf after mining, was not valid. Instead, measurements showed that the methane release was very variable, sporadically falling to zero. Carbon dioxide concentrations were monitored at the same time as methane, and followed methane concentrations. An alternative model was developed, in which methane was largely displaced from the coal and held in fissures, cleats and pores, which, when intersected by mining, released methane into the ventilation stream. In a key test of this hypothesis, methane concentrations around a continuous miner were monitored both during operation and after shutdown. Even during operation, methane became undetectable. It is hypothesized that heating of the South African coal-seams during the emplacement of the igneous dykes that cut so many of the seams caused displacement of the methane from the coal into the pores. This fissure-held methane, or free methane, is released rapidly once a pathway through the coal into the mine atmosphere is established. In contrast, methane adsorbed in the coal, and measured as the seam gas, desorbs relatively slowly. Measurements suggest that on average about 50%of the adsorbed methane is lost one day after mining. By that time, the coal will have left the mine, so that the remaining seam gas will be released to the atmosphere outside the mine. Attempts to measure methane emissions from surface coal-mining operations showed very little detectable methane. Coal from surface operations, which was milled and exposed to atmosphere, adsorbed gas from the air. It is concluded that methane release from surface operations is probably less than 3Gg/a. It is estimated that the total methane release from the coalmining industry amounts to about 72 Gg/a, as opposed to the National Communication estimates of 323 Gg/a in 1990 and 317 Gg/a in 1994. However, there are large errors associated with these estimates, and it is recommended that the methodology outlined in this report be extended to as many mines as possible to reduce the errors that are inherent in the physical processes underlying the release of methane from South African coalmines.

Jan 1, 2005

By S. Fazluddin, K. C. Sole, M. R. Serwale, T. Coetsee

The CSIR-Ti process employs lithiothermic reduction of titanium tetrachloride feedstock to produce titanium sponge. The product is therefore contaminated by a range of lithium and chloride species. In this study we examine the effects of particle size, temperature, and HCl concentration as input leaching variables on the removal of chlorides from the crude titanium sponge. A review of the aqueous chloride chemistry of Li and Ti provided initial conditions for leaching of impurity species from the sponge. Experimental results confirm that the effectiveness of leaching and removal of dissolved impurities from the sponge are dependent on leaching kinetics, which are influenced by temperature, particle size, and morphology. Of the variables tested, reaction temperature had the strongest influence on the oxygen content of the leached product. The HCl lixiviant concentration had a negligible effect under the conditions tested. Leaching of crude titanium sponge (−10 mm size fraction after crushing) at 14°C in either 1 M or 0.032 M HCl yielded a titanium sponge product that met the ASTM standard specification for commercially pure Grade 1 titanium, i.e., oxygen content < 0.18 mass% and chloride content < 0.15 mass%.

Dec 13, 2022

By W. C. Joughin

Problems with the stability and performance of rockpasses in deep-level mines are common. This paper presents the results of an investigation into the recorded performance of rockpasses in deep-level gold mines. Records relevant to the geological environment in which the passes are excavated, their excavation details, and their historical performance are not at all well documented on mines. In general, the quality of data available from mines is poor. The investigation showed that very large spans frequently developed in passes during use, and that rock mass quality, the orientation of the pass with respect to the dip of the strata, and the magnitude of the stress acting normal to the axis of the pass were the main factors of influence. From the historical information obtained, the risks of instability in rockpasses are identified.

Jan 1, 2005

By F. Peer

In most coal preparation processes, water is a necessary medium, but the presence of water in coal after it has been cleaned has a negative impact on transportation costs, handling and specific energy values. The utilization of super absorbent polymers (SAP) in the nappy application is well known. The concept of utilizing these polymers was investigated for the purposes of dewatering coal and other fines, generated by preparation processes such as flotation. SAPs are granular highly cross-linked synthetic copolymers with excellent water-absorbing properties. The dewatering process is characterized by three main stages: (a) contact of super absorbent polymer with high-moisture fine coal; (b) separation of dewatered fine coal from super absorbent polymer; and (c) regeneration of used super absorbent polymer, by exploiting its response to changes in conditions such as pH or temperature. Preliminary tests showed the separation step to be very difficult. The novel idea of encasing a given amount of polymer in a water permeable cloth solved this problem (of separating the swollen polymer from the dewatered coal). Preliminary tests investigating the effectiveness of the sachets of polymer showed a drastic decrease in the moisture contents of slurries. Furthermore, it was shown that it was possible to regenerate the polymer (still within the sachets) through thermal drying. A full-scale experimental programme was then followed to accurately determine the feasibility of using sachets of SAP. The experimental variables were initial slurry moisture content and polymer dosage. It was observed that the sachets seemed to work fastest and most effectively at high moistures and also at higher dosages of SAP. Regeneration of the polymer was also investigated, using two methods: thermal regeneration and pH-induced regeneration. The experimental variables, for investigating the regeneration process, were method of regeneration and grade of water used. Thermal regeneration at 70°C seemed to work successfully. Although the cost of using thermal energy is still a problem, the safety aspect (fire hazards) has been addressed since the polymer does not ignite when heated, unlike fine coal particles that may do so. The alternative method of regeneration, which exploits the pH-sensitivity of the polymer, was less successful, and further work needs to be conducted.

Jan 1, 2003

By J. T. Jordaan

Paper written on project work carried out in partial fulfilment of B. Eng (Mining Engineering) degree This paper addresses the critical parameters that need to be considered when designing a mechanized bord-and-pillar mining layout for a narrow tabular steep dipping orebody. These parameters are : ?Reef dip ?Reef width ?Geological structures within the orebody ?Physical dimensions of the selection of machines ?Gradeability of the selection of machines. The main objective in designing such a layout is to maximize machine efficiencies whilst minimizing dilution. This methodology was then applied in the design of a mechanized bord-and-pillar mining layout for a given piece of ground (Vlakpoort and Moddergat) using a given selection of machines. This piece of ground contains the PGM bearing UG2 Reef with an average thickness of 124 cm and dipping at 18°withadverse strike and dip changes. Two bord-and-pillar layouts were designed for this area. Layout 1 is characterized by the use of strike conveyors to convey ore on strike with an average LHD hauling distance of 35 m. Layout2 uses LHDs for all on strike conveyance of ore with an average hauling distance of 75 m. These two layouts were then technically and financially compared. It was concluded that layout 1 is theoretically superior but layout 2 was chosen due to its greater flexibility and probability of operational success. Layout 2 was also compared to conventional mining and it was concluded to be more economically feasible. However the latter is a proven mining method in these conditions and has a lower risk operational failure. Bord-and-pillar mining has some favourable promises such as improving labour difficulties, lowering operating costs and providing safer working conditions. The success of mechanized bord-and-pillar mining in narrow tabular steep dipping orebodies will depend on the extent of the research and development efforts which need to create and innovate technologies to overcome some of the current limitations stated in the above-mentioned parameters.

Jan 1, 2003

By C. Atilla Öztürk

Determining the cutability and mechanical properties of rocks is vital for the feasibility studies of a mine or tunnel projects. Many scientists have investigated the correlation between these parameters in order to predict specific energy (SE) and other important rock cutability properties for many years. On the other hand, textural properties of rocks have been quantified in detail by Howarth and Rowlands (1987). Hence, it was possible to correlate mechanical and textural properties of rocks by using a parameter known as the texture coefficient (TC). The main objective of this study was to use the test results taken from a research project carried out by Bilgin and Shahrair1tocorrelate the mechanical and cutability properties of rocks with textural properties of rocks. For this purpose, uniaxial compressive strength (UCS), uniaxial tensile strength (UTS), cone indent or (CI),point load index (PLI), unit volume weight (?), porosity, SE, and normal and cutting force values, were correlated with textural properties of rocks. Hence it was possible to make illuminating interpretations about the relationships between the mechanical and cutability with the textural properties of rocks. Keywords: Cutting properties, mechanical properties, specific energy, texture coefficient.

Jan 1, 2004

By D. P. Venter

When Oryx Mine down-scaled in 1994, focus was placed on development, with the primary aim being the exposure and proving of ore reserves. In order to achieve this aim, development ends had to be advanced to reef at the maximum possible rate. This in turn called for support methods which could keep up with the advancing faces, taking up minimum space and time for installation, while providing on the face a real coverage and effective long-term support. In the majority of the development ends which were located in fairly competent ground, the grouted rod development support proved to be adequate until such time as secondary support in the form of meshing and lacing could be installed. The ends developing towards the eastern boundary of the mine, especially in the ?C? block, encountered extremely adverse ground conditions due to the presence of smectite zones within the footwall quartzites. Smectites are a group of montmorillonite clay materials, mainly composed of hydrous aluminosilicates. They occur as minute platy crystals in the rock mass and are especially notable for their ability to absorb liquids. Generally, the zones are found in between the bedding planes of the host rock, and may vary considerably in thickness over short distances. Wherever the smectites were exposed by the development blasting, the rockwalls weathered as the smectites absorbed water and the condition of the tunnel rockwalls deteriorated rapidly. When the smectites were left exposed and allowed to weather, it became almost impossible to install secondary permanent support such as meshing and lacing at a later stage. To minimise the exposure time and subsequent deterioration of the smectite zones, it was recommended that all development ends encountering smectites be shotcreted as soon as possible after blasting. In developing a shotcreting support methodology, both a drycreting and wetcreting system were investigated. The situation at the time ruled out the use of a wetcreting system on the basis of the extensive and costly infrastructure required, as well as its limited manoeuverability. Instead, it was decided to opt for as maller scale but more manoeuverable drycreting system. The system utilized a number of portable drycrete machines with pre-bagged cement, which could support several development ends in close proximity. Initial trials using normal 30 MPa shotcrete on 19level proved successful and the system was implemented.

Jan 1, 1998

By N. D. L. Burger

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 steel wire rope 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 rope. This detaching is carried out via an additional detaching 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 to conform to 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 a1:10 scale shaft model. Retardation performance close to the required levels was achieved. The retardation distance required should allow such systems to be retrofitted in most existing shafts. Given the simplicity and robustness of the designs, further development was recommended because of their enhanced self-sufficiency and reduced risk of malfunction. Keywords: underwind, overwind, deceleration, mineshafts, hoisting.

Jan 1, 2004