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By E. Van Der Lingen

Within the context of the overall gold market industrial demand plays an important role. Industrial demand for gold was 430 tonnes in 2008, approximately 12% of total end-user consumption. The dominant consuming industry in this market for gold remains the electronics sector, consuming an estimated 300 tonnes, primarily through the use of gold bonding wire for computer chips and plated contacts and connectors. Substantial demand also derives from dental applications, with other smaller industrial markets and decorative uses of the metal (non-jewellery) constituting the balance. In each market gold faces tough competition from alternative lower cost materials, such as silver and copper (electronics) and ceramics (dental). As well as the obvious benefits to the gold market resulting from this constant physical demand, there are an umber of additional positive factors concerning the use of gold in industrial applications: ?Price inelasticity: demand for gold from the major consuming industries has so far proven to be relatively robust in the face of significant increases in gold price. ??Stickiness?: the gold consumed by industrial uses tends not to return easily to the market (i.e. the demand is ?sticky?) because it is highly dispersed. ?Image: the beneficial environmental and medical applications of gold (e.g. gold used in the treatment of prostate cancer and in purification of contaminated water) are helpful to the positive image of gold. ?Added-value: new industrial uses for gold can create substantial added value in the supply chain, particularly in gold producing countries.

Jan 1, 2009

Limestone formations in the United States can be subject to relatively high horizontal stresses due to the existence of tectonic loading of the limestone strata. Underground limestone mines use the room-and-pillar method, in which 12- to 18-m-wide rooms are typically excavated. The stability of these excavations can be compromised by the horizontal stress, resulting in a rockfall hazard. Rockfalls are the cause of 15% of all reportable injuries in underground limestone mines. Horizontal stress related damage can occur in the form of guttering along one or more sides of an excavation, roof beam buckling or oval shaped roof falls, with the long axis perpendicular to the major horizontal stress. Numerical analyses show that the pillar layout and orientation of the mine workings have an effect on the horizontal stress distribution within the roof. The effects of high horizontal stresses can be mitigated by orientating the heading development direction parallel to the maximum horizontal stress, reducing the number of cross-cuts and offsetting the cross-cuts to limit the potential lateral extent of horizontal stress related roof falls. The modeling approach described in this paper can be used as a tool to evaluate potential roof failure and optimize the stability of room and pillar layouts.

Jan 1, 2008

By H. Gumede

The challenges of noise emission in the mining industry are well known. These challenges are exacerbated by the increasing trend towards mechanized mining, and there is general consensus that an effective industry-wide Buy Quiet initiative (BQI) will significantly assist in reducing noise-induced hearing loss by minimizing noise at its source. In its endeavours to improve the health of workers, the MOSH Noise Team has facilitated development of a proposed strategy and structure for a BQI. Extensive consultation with stakeholders resulted in support for the initiative. The MOSH Noise Team was advised to solicit the contributions of consulting mechanical and electrical engineers (CM&EEs), whose support is viewed as critical for realization of the initiative, since elimination of noise at source is an engineering challenge, and most procurement standards and specifications are set and managed by engineers. A workshop with CM&EEs and their representatives, as well as group environmental engineers (GEEs), confirmed the urgent need for the initiative and provided valuable insights. The workshop used obstacle based planning techniques to identify the factors on which the success of this BQI will depend. Most potential obstacles to the implementation revolved around proper scoping, leadership commitment, and companies? adherence to the initiative. Lack of involvement by the relevant stakeholders and economic issues were also identified as potential obstacles. Delegates at the workshop felt that the successful implementation of the industry-wide BQI, will depend on the mining industry addressing the specific obstacles. The most critical and urgent issue was identified as the formation of a properly constituted industry-wide BQI task team that includes representatives from critical stakeholder groups and which is mandated and empowered to drive the initiative.

Jan 1, 2014

By J. O. Marsden

The copper industry is cyclical with the price determined by market forces (i.e. the balance between supply and demand for the metal.) The period of each cycle typically varies between 6 to 9 years. These cycles drive decision making for exploration, new project and mine development, mine closures, process improvement, research and development, and technological innovation. The gold industry is fundamentally different from copper in that supply and demand do not, alone, determine metal price and many other factors play a key role, including global and regional financial conditions, political stability, and the global future economic outlook. However, the gold price experiences trends and longer term cycles that similarly drive business decision making. This paper examines developments in the copper industry that have occurred over the past few price cycles (approximately 20 years) and explores how the lessons learned can be applied to gold extraction. A number of specific examples and opportunities are discussed in detail, including; technology development and innovation, cost control and business improvement, material characterization, comminution equipment, biological heap leaching, and concentrate pressure leaching.

Jan 1, 2009

By G. Dhakal, J. -I. Kodama, K. Neaupane, T. Yoneda

ABSTRACT Durability tests and evaluation of sub-zero temperature effects to some argillaceous clastic rocks were carried out. Welded tuff, limestone, sandstone and schist samples collected from different parts of Japan were tested and results are presented. Schist and dolomite from Ganesh Himal area along the higher Himalayas of Nepal were also tested, and the results are compared. The slaking tests were performed on all samples in pure water. The results reveal that the slaking index of schist from Japan is lower than that of schist from Nepal. Regarding carbonate rocks, the slaking index of limestone from Japan is higher than that of dolomite from Nepal. A tendency of decrease in slake durability with an increase in porosity is also observed. Freezing and thawing experiments conducted on welded tuff and schist reveal that crack initiates earlier but extends slower in schist than in welded tuff. It is also observed from the freeze-thaw experiment that the welded tuff deteriorates and wears faster than the schist. However, sandstone remains relatively intact for a long run of freeze-thaw cycles. Zusammenfassung Beständigkeitstests sowie die Bewertung der Auswirkungen von Minusgraden auf Tonschiefer Trümmergestein wurden durchgeführt. Dabei wurden geschweißter Tuff, Kalkstein, Sandstein und Schiefer aus verschiedenen Gebieten Japans sowie Schiefer und Dolomitgestein aus dem Ganesh Himal Gebiet, dem höheren Himalaya Nepals, untersucht und die Resultate verglichen. Für alle diese Gesteine wurden Slaking-Tests in reinem Wasser vorgenommen. Die Resultate zeigen, dass das Slaking von Schiefer aus Japan höher als das von Schiefer aus Nepal war, wohingegen bei kalkhaltigem Gestein das Slaking von Kalkstein aus Japan niedriger als das von Dolomitgestein aus Nepal ausfiel. Es wurde dabei auch eine Tendenz zur Löschbeständigkeitsverringerung bei gleichzeitig erhöhter Porosität beobachtet. Gefrier- und Tautests an geschweißtem Tuff und japanischem Schiefer zeigten, dass Risse in Schiefer schneller auftraten, aber sich langsamer ausweiteten als in geschweißtem Tuff. Bei Gefrier- und Tautests wurde ebenfalls beobachtet, dass geschweißter Tuff schneller verfällt und verschleißt als Schiefer. Sandstein hingegen blieb über lange Tau- und Gefrierzyklen intakt Résumé Des tests de solidités et l’évaluation de l’effet des températures en dessous de zéro sur les rocs argileux ont été effectués. Les rocs d’origine volcanique (cendre), les rocs calcaires, sableux et les schistes de différentes régions du Japon, ont été étudiés. Les schistes et les dolomites de la région de Ganesh Himal le long du haut Himalaya (Népal) ont été étudiés et comparés à ceux du Japon. Les tests de désintégration sur les rocs ont été effectués dans l’eau pure. Les résultats montrent que l’index de désintégration des schistes du Japon est inférieur aux schistes du Népal. En ce qui concerne les rocs carbonyles, l’index de désintégration des rocs calcaires du Japon est supérieur à ceux des dolomites du Népal. Une décroissance de la résistance à la durabilité est observée avec une plus grande porosité des rocs. Les tests de congélation-décongelation sur les schistes et les rocs d’origine volcanique (cendre) du Japon montrent une apparition précoce des fissures mais l’expansion des fissures est plus rapide dans les rocs volcaniques. Des tests similaires montrent que les rocs volcaniques (cendre) se désagrégent plus rapidement que les schistes. Cependant, les rocs sableux restent intacts après plusieurs cycles de congélation-décongélation.

Jan 1, 2003

The Namakwa Sands heavy mineral deposit is located along the West Coast of South Africa and the mine is a world class producer of high quality zircon, ilmenite and rutile concentrates from essentially unconsolidated marine and aeolian sands of Cainozoic age. The objective of this study was to characterize rutile with respect to distribution, grain size, textures, colour and mineral chemistry within the Namakwa Sands orebody with the aim to explain its overall poor recovery. A representative suite of heavy mineral concentrates from various sections of the orebody and stages in the recovery circuit has been investigated microscopically and by means of SEM-EDS and LA-ICP-MS. The rutile grain size distribution displays a wide range and is directly related to its variable chemistry and consequently density. In placer deposits such as Namakwa Sands, mineral sorting is a function of hydraulic equivalence, and high density grains are smaller than the lighter grains when deposited under similar conditions. Grain size is also a function of sediment maturity, and the highly mature red aeolians and (RAS) component of the deposit has a coarser grain size than the less mature orange feldspathic marine sands (OFS). The above primary characteristics of rutile are responsible for the loss of the coarse-grained fraction during screening and, given the empirically determined relationship that increased substitution elements reduce conductivity, also during electrostatic separation. Provenance studies using geothermometry have shown that the heavy mineral suite has been sourced mainly by the proximal medium-to high-grade Namaqualand Metamorphic Complex. The positive relationship between substitution elements and temperature of formation explains their high concentration in rutile of this deposit. The heterogeneity of rutile, produced by the combination of a high temperature primary source and the typical marine-aeolian placer genesis is deposit specific and unfortunately not conducive to high recovery levels in the current Namakwa Sands beneficiation circuit.

Jan 1, 2010

Geologists in some underground gold mines collect grab samples from broken ore piles or trucks as a method of grade control. It is often known as muck sampling. Generally, the goal of grab sampling is to try and reconcile the mined grade at the ore source to the predicted grade and/or predict the mill feed grade. The mass of the sample collected is limited by health and safety issues, as well as the capacity of the laboratory to process the samples within a given timeframe. In general terms, grab sampling is known to be problematic because samplers tend to over sample the fines, and/or pick out high-grade fragments; surface sampling of piles does not test material within the pile; muck piles in development drives/faces are likely to be zoned due to the blasting sequence; high or low-grade material may preferentially segregate in the pile during mucking; the five percent mass reject size of the material in muck piles is very large from underground blasting; some correlation usually exists whereby the larger fragments are enriched or depleted in the critical component of value; and the average error made in estimating the true stockpile grade is likely to be high. The method is prone to chronic fundamental sampling, grouping and segregation, delimitation, and extraction errors. Substantial warnings must be given with respect to the use of grab sampling for grade control in gold mines. The method may appear to work sometimes, which can be attributed to a fine gold particle sizing and more disseminated distribution. As with all sampling methods, its appropriateness must be determined by ore characterization and heterogeneity testing to ensure the method suits the ore type.

Jan 1, 2009

What is a nugget effect? In the early development of geostatistics, the term ?nugget effect? was coined for the apparent discontinuity at the beginning of many semi-variogram graphs. This name was chosen to reflect the large differences found between neighbouring samples in ?nuggety? mineralizations such as Wits gold reefs. Geostatistical theory assumes that the difference between a sampled value and a potential repeat sample at the same location is actually zero. Included in this ?nugget effect? would be true variation between contiguous samples due to the nature of the mineralization, micro-fracturing, nugget or crystal size and so on. Also included in the nugget effect would be any ?random? sampling variation which might occur due to the method in which the sample was taken, the adequacy of the sample size, the assaying process, etc. Arguments were put forward that ?sampling errors? actually exist at zero distance. Some geostatistical schools actually maintain that ?nugget effect? is all sampling error. This would imply that ?perfect? sampling would eliminate the nugget effect entirely. There is now a dichotomy both in the geostatistical world and in the software packages provided for geostatistical analyses. It may seem academic to argue over whether the semi-variogram model should take a value of zero, a value equal to the nugget effect or a partial value at distance zero. However, the decision can have a profound effect on both the estimated resource and in our confidence on that resource. Whereas most geostatistical texts define the semi-variogram model as taking the value of zero at zero distance, others imply that the full nugget effect should be used at zero distance. For example: ?The nugget effect refers to the nonzero intercept of the variogram and is an overall estimate of error caused by measurement inaccuracy and environmental variability occurring at fine enough scales to be unresolved by the sampling interval3. ?Christensen4 has shown that the ?nugget effect?, or non-zero variance at the origin of the sernivariogram, can be reproduced by a measurement error model. ?The nugget effect is considered random noise and may represent short-scale variability, measurement error, sample rate, etc.5. In many training texts and web courses, the definition of the semi-variogram is ambiguous as the formulae for semi-variogram models is not actually specified at zero distance6,7,8.

Jan 1, 2009

This paper examines manning best practice in the Australian mining industry, comparing the mining of gold with mining of other commodities. In particular, the paper considers a range of mining operations and their manning levels as measured by operational benchmarking. The benchmarking results presented cover 7 open pit and 14 underground studies of gold mines in Australia from a database of over 50 studies. The data were collected as part of AMC?s benchmarking and operational improvement studies conducted between 1999 and 2009. The results examine overall manning levels in mining, processing and administration functions. Manning levels for a range of mining methods, including open pit and underground methods, are segmented. Productivity measures such as ore tonnes per man day and metres developed per man day are presented and where direct relationships can be established, the relationships are described. The proportion of supervisory to operations personnel and labour costs as a percentage of total operating costs is highlighted. Minimum manning requirements for the range of mining methods are indicated, irrespective of the size of the operation. The benchmarking results provide guidance by indicating a range of manning numbers and trends which are likely to be appropriate for the scale and type for a range of mining operations. When combined with benchmarked productivity and cost data, manning results help identify potential areas for business improvement, driving gains in the efficiency of labour. Finally, the paper considers perspectives on best practice for manning and the limitations in application of this type of analysis in a global context.

Jan 1, 2009

Good blasting practices based on the careful application of explosive energy lead to significantly safer mining operations. Case studies demonstrate how the selection of the correct charge mass, explosive type and round design are vital for extracting the exact amount of rock required and preventing spurious damage to the remaining rock mass. In other cases the explosives are applied to precondition the rock mass, design rockburst resistant support, prevent damage to important excavations and reduce the exposure of miners to unsafe conditions. The Hybrid Stress Blasting Model is used to understand the explosive rock interaction and to illustrate and contrast good and bad blasting practices. In all cases, good supervision is required to continue safe mining according to the design. Keywords Blasting, safety, production, preconditioning, explosives, numerical modelling, drop raising, vibrations, and presplitting.

Jan 1, 2011

By S. Tu, Y. Yuan, H. Wei, Z. Chen, M. Jia

"To reduce production costs and increase efficiency in ageing coal mines, a system that utilizes two shearers in a fully mechanized longwall working face is proposed. Using theoretical and engineering experience, three issues pertaining to the system were determined: (a) the two-shearer mining technique; (b) matching and modifying the equipment; (c) and coal-cutting task allocation for each shearer. Two mining processes are proposed, with the two shearers travelling in either the same or opposing directions. To avoid breakage of chains and pan extrusion of the armoured face conveyor (AFC), the length of the AFC ‘snake’ was controlled. To ensure continuous cutting and transporting of coal, the cross-sectional dimensions of the AFC were checked and the shearer closer to the head drive was modified. Based on the assumption of equal cutting times for each shearer, the meeting position of the two shearers was obtained by using a theoretical model of mining task allocation. Application of this technique at No. 2 Jining Mine, China, has shown remarkable benefits: daily production capacity was increased by 54% and personnel efficiency was improved by 33 t/d per person. IntroductionAll coal mines aim to increase the unit output of the working face to reduce mining costs and improve the economic efficiency of the operation. To achieve this goal, a production model known as one mining face of the mining area has been adopted by most Chinese coal mines (Hu, Meng, and Zhu, 2008; Zhang, Zhang, and Wang, 2000). Under these conditions, the main technical approaches to maximize the output are increasing the width of the working face (Qu, Xu, and Xue, 2009) and accelerating the advancing speed (Robbins, 2000). In coalfields with shallow seams, the width of a fully mechanized face exceeds 300 m (Ju and Zhu, 2015; Fu, Song, and Xing, 2010); in coalfields with deep seams, the width of the face is usually greater than 240 m (Liu et al., 2016; Li et al., 2013). The term ‘super-long working face’ was proposed to describe working faces with a width of over 240 m (Zhao and Song, 2016; Xu et al, 2007) in China. To support these long working faces with rapid advancing speeds, high-powered mining equipment is required (Kulshreshtha and Parikh, 2001, 2002; Tu et al., 2009; Mishra, Sugla,and Singha, 2013). This is easily achieved in newly built mines, but is not a good choice for ageing mines because of the low return on a high investment on account of the limited remaining resources. It is therefore difficult for ageing mines to significantly increase their unit output with the existing mining equipment. This study focused on this problem: two shearers were applied to a longwall fully mechanized working face (LFMWF) to achieve increased unit output."

Oct 1, 2018

By J. P. Harrison, J. A. Hudson.

ABSTRACT: When considering the rock stress component of rock mechanics design, the basic procedure is to consider the pre-existing stress-state in the rock mass and then estimate the disturbance to this stress state as a result of engineering excavation. Failure criteria are applied to the final stress state to establish whether failure will occur. However, during the evolution of the stress state, the failure locus may be reached before the anticipated failure state and, in this case, failure will occur unexpectedly. This indicates that the complete stress path (defined as the variations in magnitudes and orientations of the stress tensor components resulting from engineering-induced or natural changes) must be considered. We list a set of stress path principles with their corollaries and then introduce a new and powerful method for simultaneously illustrating changes in the six independent components of the stress tensor. This method plots the magnitudes of the principal stresses above their orientational paths on the hemispherical projection, to provide a three-dimensional stress path locus for each of the three principal stress components. In connection with this presentational device, we illustrate a number of concepts associated with the stress path in rock engineering. For example, we show how the minor principal stress approaches zero at an unsupported excavation face, the interchange of the trajectories of two principal stresses when their magnitudes become equal, and the influence of the bounds on the principal stresses given by strength criteria. RESUME: Si on s’interesse à la composante “contrainte” lors d’un dimensionnement en mécanique des roches, la méthode de base consiste à considérer l’état de contrainte pré-existant dans le massif rocheux, puis d’estimer la perturbation à cet état générée par l’excavation projetée. Des critères de rupture sont appliqués à l’état de contrainte final pour établir si une rupture va se produire. Cependant, lors de l’évolution de l’état de contrainte, la surface critique peut être atteinte ce qui provoquera une rupture anticipée et, dans ce cas, cette dernière sera inattendue. Cela indique que le chemin de contrainte complet (défini par les variations de l’intensité et de l’orientation des composantes du tenseur des contraintes liées à des causes anthropiques ou naturelles) doit être pris en compte. Nous présentons un ensemble de règles pour les chemins de contraintes ainsi que leurs corollaires et, ensuite, nous introduisons une méthode, nouvelle et puissante, pour présenter simultannement les changements des six composantes du tenseur des contraintes. Cette méthode consiste à tracer l’intensité des contraintes principales sur le chemin d’orientation des contraintes à l’aide d’une projection hémisphérique afin d’obtenir le lieu, en trois dimensions, de chaque contrainte principale. En relation avec cet outil de représentation, nous illustrons certains concepts associés au chemin de contrainte en ingénierie au rocher. Par exemple, nous montrons comment la contrainte mineure tend vers zero à la surface d’une excavation sans soutenement, le changement de direction des deux contraintes principales quand leur intensité deviennent égales, et l’influence des conditions aux limites sur les contraintes fournies par un critère de rupture. ZUSAMMENFASSUNG: Bei ingenieurmassigen Taetigkeiten in Felsgesteinen ist es noetig die ungestoerten Gesteins Stress Componenten, die vor einem anthropogen Eingriff im Gestein vorhanden waren, zu kennen und erst dann Vorhersagen ueber Stoerung des Stress Zustandes, der durch einen ingenieurmaessigen Eingriff hervorgerufen wird, herzuleiten. Kriterien fuer das Versagen werden im letzten Stadium, des Stress Zustandes angewendet, um zu erfahren ob ein Versagen wohl eintreten wird. Es ist jedoch zu beachten, dass waehrend des Stress Aufbaus ein oertliches Versagen eintreten kann, welches vor dem eigentlichen, erwarteten Versagens Zustand auftritt. Diese Moeglichkeit weist darauf hin, dass der gesammte “Stress Path” (definiert durch die Variationen der Stress Tensor Komponenten in Richtung und Groesse, hervorgerufen durch ingenieurmaessig hervorgerufene oder natuerlich Vorgaenge) betrachtet werden muss. Wir fuehren eine Liste der “Stress Path” Prinzipien auf und stellen dann eine neue und wirksame Methode zur gleichzeitigen Darstellung der sechs unabhaegigen Komponenten des Stress Tensors vor. Diese Methode stellt die Groesse der “Principal Stresses” in Abhaengigkeit von ihrem rauemlichen Weg auf einer hemispherischen Projektion dar und erlaubt hiermit eine drei dimensionale Darstellung fuer jeden Ort des “Stress Path’s” in die drei Haupt Stress Komponenten. Im Zusammenhang mit der hier gegebenen Darstellungsart, verdeutlichen wir eine Reihe von Konzepten in der Felsmechanik. Wir zeigen zum Beispiel, wie der “Minor Principal Stress” Null zustrebt an einer Wand von einer nicht abgestuetzten Ausschachtung, den Austausch von Trajektorien zweier “Principal Stresses” wenn ihre Groessen sich gleichen, und den Einfluss der Groesse von “Principal Stresses” bei gegebenen Kriterien des Festgkeits Versagens.

Jan 1, 2003

By C. Musingwini, C. Beaumont

"Kansanshi mine, located in northwestern Zambia, uses conventional openpit mining to produce primarily copper, with gold as a by-product. The ore is extracted from a deposit with highly complex mineralogical suites that are difficult to process, sometimes leading to poor recoveries. The mine was using an ore classification system called ‘Mat-Type’, which classified various ore types into 22 different quality categories, ranging from poor to good quality. Ore with poor recovery was classified as ’poor quality’ ore and directed to long-term stockpiles. Due to a lack of proper integration between the technical and functional areas forming the mine value chain, Mat-Type unintentionally misclassified ore. A geometallurgical analysis was therefore undertaken to review Mat-Type and a new geometallurgical ore classification system, called ‘OXMAT ’, was developed to replace Mat- Type.When implemented, OXMAT eliminated nine of the 22 ore categories and the remaining 13 ore categories could now be more accurately and consistently defined, leading to a significant change in the quantity of ore being mined. OXMAT demonstrated that within a defined test volume of the geological model, waste tonnes could be reduced by 19% while ore tonnes could be increased by 17%. This paper therefore demonstrates the value created through integrated geometallurgical modelling. IntroductionKansanshi mine is located in northwestern Zambia, approximately 10 km north of the provincial capital Solwezi and about 160 km west of Nchanga mine in the Central African Copperbelt (Broughton and Hitzman, 2002). The location of Kansanshi mine relative to other major copper deposits in the Copperbelt is indicated in Figure 1.Kansanshi mine uses conventional openpit mining methods to exploit copper- and gold-bearing veins hosted in a structurallycontrolled mineral deposit. Mining is carried out in two areas of the deposit, the Northwest and Main pits. A third area of the deposit, called Southeast Dome, is currently not being exploited (Figure 2)."

Mar 1, 2019

By T. Feroze, B. Genc

"Blind headings in room and pillar coal mines are a major source of methane and coal dust. Most methane and coal dust explosions therefore occur in the blind headings. The primary cause of these explosions is the disruption of the local ventilation system. Line brattice (LB) ventilation systems are used to ventilate blind headings by directing air from the last through road (LTR) into the heading. The amount of air available to ventilate the face of the empty heading depends on the heading dimension, settings of the LB, and velocity of air in the LTR. LBs are commonly installed by underground supervisory staff based on work experience, which may result in ineffective ventilation. The correct installationof LBs remains a challenge. In this study, a validated computational fluid dynamics (CFD) model has been used to analyse the effect of the LB ventilation system variables on the air flow rates close to the face of the empty heading. Full-scale three-dimensional models with various heading heights, heading depths, LB settings, and LTR velocities were simulated. The air flow rates and patterns at various locations inside the heading were analysed. A comparative study was carried out to quantify the effect of each of these system variables on the flow rates close to the face of the heading. Based on the findings, a user-friendly numerical model was formulated that can be used to estimate the flow rate close to the face of an empty heading for different practical settings of the system variables. This model can help the supervisory staff to swiftly implement the ventilation plan according to the regulations and the mine standards. The model can also serve as part of the curriculum for educating future mining engineers."

Jan 1, 2017

By T. Feroze, B. Genc

"Blind headings in room and pillar coal mines are a major source of methane and coal dust. Most methane and coal dust explosions therefore occur in the blind headings. The primary cause of these explosions is the disruption of the local ventilation system. Line brattice (LB) ventilation systems are used to ventilate blind headings by directing air from the last through road (LTR) into the heading. The amount of air available to ventilate the face of the empty heading depends on the heading dimension, settings of the LB, and velocity of air in the LTR. LBs are commonly installed by underground supervisory staff based on work experience, which may result in ineffective ventilation. The correct installation of LBs remains a challenge.In this study, a validated computational fluid dynamics (CFD) model has been used to analyse the effect of the LB ventilation system variables on the air flow rates close to the face of the empty heading. Full-scale three-dimensional models with various heading heights, heading depths, LB settings, and LTR velocities were simulated. The air flow rates and patterns at various locations inside the heading were analysed. A comparative study was carried out to quantify the effect of each of these system variables on the flow rates close to the face of the heading. Based on the findings, a user-friendly numerical model was formulated that can be used to estimate the flow rate close to the face of an empty heading for different practical settings of the system variables. This model can help the supervisory staff to swiftly implement the ventilation plan according to the regulations and the mine standards. The model can also serve as part of the curriculum for educating future mining engineers."

Jan 1, 2017

By M. D. Raganya

The Kriel Colliery is an Anglo Coal mine which is contracted with Eskom to produce No. 4 seam for Kriel power station. The mine previously found it uneconomical to mine No. 5 seam due to the low qualities and tonnages of the seam, both at the opencast and underground sections. An area has been encountered for the first time at Kriel Colliery opencast pit 5, where the tonnage and qualities of No. 5 seam warrant further investigation. The primary reason for the investi¬gation of No. 5 seam by Kriel Colliery is because of the advantage of the high selling price (R450/t) that the market is currently prepared to pay for the No. 5 seam. The aim of this project is to mine No. 5 seam economically without interrupting No. 4 seam mining operations. The problem is that the equipments that are currently being used at the mine belong to Eskom and are dedicated to mining No. 4 seam only, not No. 5 seam. Through geology investigation, it is found that the whole No. 5 seam (No. 5 seam upper and lower) has an average thickness of 1.8 m. It also contains a parting of carbonaceous shale which divides the seam into two parts. No. 5 seam upper contains less tonnage compared to No. 5 seam lower. The calorific value and ash content of No. 5 seam lower is far better than No. 5 seam upper, but would still require washing the coal. This makes No. 5 seam lower the preferred seam to mine because it meets the requirements of the customer which is Highveld Steel. It is a coincidence that the area containing No. 5 seam under investigation encounters high overburden problems. The problem of drilling to No. 4 seam was going to be encountered because the thickness of the overburden ranges from 35 to 39 m. The Gardner Denver at the mine can drill blastholes only to a depth of 34.5 m, meaning the full overburden will not be drilled at once. This means the mining of No. 5 seam will do favours for Eskom because the overburden thickness will be minimized. This then led to an agreement between Anglo Coal and Eskom to mine No. 5 seam because both companies benefit from the project. In order to identify a suitable mining method for No. 5 seam lower, literature studies took place. It is found that San Juan mine from the USA has similar geological features to Kriel Colliery's No. 5 seam. Through comparisons, it is found that truck and shovel mining would suit the geology of Kriel Colliery's No. 5 seam since San Juan mine is using it. The sequence of mining No. 5 seam lower allows for a clearance of two cuts between the truck and shovel operation and the dragline. The mining sequence of No. 5 and 4 seam will be in the order of top soil removal; soft overburden pre-stripping; pre-strip drilling; pre-strip hard overburden; No. 5 seam lower coaling; interburden drilling and blasting; interburden stripping and No. 4 seam coaling. Financial evaluation of the project shows that a contractor will charge Kriel Colliery (Anglo Coal) approximately R 102.2 million to mine No. 5 seam successfully. The sales of No. 5 seam lower will provide a profit of R99.4 million at a yield of 70% from the washing plant. The project is beneficial to both Eskom and Anglo Coal because Eskom will have eliminated the problem of pre-stripping and Anglo Coal will generate additional profit. Through investigation and analyses of the project, it is proven to be viable and should take place. This is after looking at the advantages that the project has for Anglo Coal and Eskom. Keywords No. 5 seam, truck and shovel mining, Kriel Colliery, No. 5 seam mining at Kriel Colliery opencast, Anglo Coal and Eskom, No. 5 seam upper and lower, top soil removal, soft overburden pre-stripping, pre-strip drilling, pre-strip hard overburden, No. 5 seam lower coaling, interburden drilling and blasting, interburden stripping, No. 4 seam coaling.

Jan 1, 2011

By J. A. Theron

"This paper provides guidelines on performing mass and energy balance modelling involving coal and coal derivatives. Usually, the inputs to a pyrometallurgical process would be specified in terms of elements and compounds. Reliable thermochemical data is more widely available for species involving uniquely defined, relatively smaller molecules. However, in the case of coal, the molecules are extremely large and not uniquely defined. Consequently, modelling processes involving coal and its derivatives have several potential pitfalls. These are outlined in the present paper. The main findings are as follows:•,Coal proximate analysis should not be regarded as absolute; it could vary with several parameters, including heating rate •,For modelling, the use of ultimate analyses should be considered a preferable option to proximate analyses, where, ‘fixed carbon’ and ‘volatiles’ are not defined in terms of chemical composition •,Significant errors could be made if the larger molecules are neglected during calculation of the calorific value (CV) of coal gas (the gas liberated when coal is heated in the absence of oxygen) •,For elemental analyses, the oxygen content (which is calculated by balance) should be checked to ensure it is within the expected range •,For representation of sulphur in coal, double-counting due to SO3 in the ash analysis should be avoided •,Potentially, oxygen in coal could be represented as O2, H2O, CO, or CO2; however, use of some of these species without considering the experimentally determined gross CV results in significant errors in the energy balance. If coal enthalpy is calculated from elemental analyses without correction, representation of coal oxygen as H2O(l) gives reasonable accuracy (up to 6.9% error for a set of South African coals). Coal volatiles could be represented by a complex mixture of compounds, even using different oxygen-containing species than the abovementioned four, provided the enthalpy is corrected •,It is recommended that an ‘enthalpy correction value’ is incorporated in energy balances involving combustion, devolatilization, or conversion of coal and coal derivatives, e.g. coke, char, or tar. That would imply that proximate analysis, elemental analysis, as well as the gross CV would be required for all solid or liquid coal-derived substances being modelled. No other correction due to the form of carbon being different than graphite should be used, as that would imply doublecounting some effects."

Jan 1, 2014

By S. R. Grobler, S. Bilessuris, M. E. Greenway

Ensuring that all the head ropes on a multi-rope friction winder share the load equally has been a design and maintenance issue since they were first developed. Accurate machining of the rope grooves is necessary to meet this objective. Early analysis of the change of rope loads for ropes in grooves that are mismatched in depth treated the grooves as rigid. Groove depth tolerances deduced from this analysis are stringent and difficult to achieve in practice. Test work indicated that the inherent flexibility of groove lining materials alleviates the load-sharing problem. However, no analysis of rope load changes due to mismatched flexible grooves has been published. In this paper we develop new equations for the rope load variation for flexible grooves. By allowing flexibility to reduce to zero, the corresponding rigid rope groove equations are obtained. New criteria for tolerable groove depth variations are developed from this analysis, which depend on the flexibility of the groove lining material.

Sep 1, 2021

By B. K. Loveday

A Colloquium on 'Control and Simulation in the Minerals Industry' was held at the Holiday Inn, Milner Park, on 20th September, 1984, in conjunction with the Electra Mining Exhibition. A total of 170 persons attended the very full morning programme, which began with an opening address by the President, The South African Institute of Mining and Metallurgy, Mr l.D. Austin. He outlined the reasons why computers were being used to an ever-increasing extent for control and simulation in the minerals industry. The computer was even used to check, among other things, the validity of the signals it was receiving by mass-balance calculations.

Jan 1, 1985

By M. Paricheh, M. Rahmanpour

"The cost efficiency and high reliability of semi-mobile or combined in-pit crushing-conveying (IPCC) and truck systems, compared to conventional truck-shovel systems alone, make them more attractive for use in modern mining operations. Semi-mobile systems, which are the most common systems in open-pit mining, combine the advantages of both systems such that fewer trucks are required, operating costs are lowered, and the operation is environmentally sound. One major aspect of utilizing this system is determining the location of the in-pit crusher and the timing of relocations. Facility location models are rarely used in the mining sector, but these models may be used for solving the in-pit crusher location problem. To this end, the main parameters affecting the IPCC location in open pit mines are reviewed and examples given of how they are applied to a dynamic location-relocation problem. Subsequently, the model is implemented for Sungun copper mine in Iran and the number and the exact time of relocations of the IPCC units determined. IntroductionNowadays, compared to the last century, metal prices are more volatile, average and cut-off grades have decreased, and stripping ratios increased. From a mining engineering point of view, many ore deposits with simple access and low capital and operating costs have been exhausted. The remaining deposits are located at depth, with no easy access, which requires high initial capital costs (Osanloo, 2012). Mining at great depths and high tonnages is possible only through open pit mining. Today, more than 80% of open pit mines in the world use a shovel-truck system for loading-hauling operations (Osanloo, 2010).With increasing depth, haulage distances increase and the number of loads per truck decreases. As a result, fuel, tyre, and depreciation costs per ton increase. Owing to these factors, open pit mining at great depth (300 to 1000 m) using shovel-truck systems faces some technical and economic problems. In-pit crushing-conveying (IPCC) systems have been known in the mining industry for many decades. The idea was introduced in 1956 in Germany. Today, the cost efficiency and high reliability of IPCC make it more appealing than conventional shovel-truck operation, especially in longer life projects with lengthy transportation distances and high production rates (Koehler, 2003). Since 1956, different aspects of this system have been reviewed by many researchers. Barua and Lanergan (1985) developed a computer program that compares the tonnage aspects of various conveyor layouts. They believed that to maximize the economic benefit of lower haulage costs associated with in-pit conveying, one must first minimize the cost of the flatter slope, which is expressed in terms of additional stripping or tied-up ore. Dos Santos and Stanisic (1987) evaluated the design of a high-angle conveyor at the Majdanpek copper mine. Then, Sturgul (1987) applied GPSS (general purpose simulation system) to find the best location of an in-pit crusher."

Jan 1, 2017