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By P. F. Ledwaba

"Artisanal and small-scale mining (ASM) in South Africa received official recognition after the change in government in 1994.The Reconstruction and Development Programme (RDP) recognized the sector as a vehicle for social and economic development for historically disadvantaged South Africans (HDSAs) who had previously been excluded from participating in the mainstream economy. Having recognized the ASM sector, government introduced several interventions and support structures to foster the development of ASM, to encourage participation of HDSAs, and to address challenges facing the sector. The objective of this paper is to assess the progress made, with particular regard to challenges in the sector. The paper focuses on the policy requirements that were deemed important for the growth of the ASM sector by the White Paper on Minerals and Mining Policy (1998). These can be collated into five categories: access to mineral rights, access to finance, access to markets, technology and skills, and institutional support. The paper provides a review of the support interventions, their intended roles and impact on the sector, and identifies existing gaps and possible ways of dealing with the challenges. There is a need for research to assess the real impact of these past and existing interventions on the ASM sector to draw lessons for future development.IntroductionThe significant rise in artisanal and smallscale mining (ASM) activities worldwide has led to many countries recognizing the sector. ASM takes place in approximately 80 countries (World Bank, 2013). It is estimated that the sector employs between 20 and 30 million people around the world (Buxton, 2013). This compares to between 3 and 3.7 million in 1999 (International Labour Organization, 1999). ASM activities are widespread, occurring mostly in developing countries in Africa, Asia, Oceania, and Central and South America (World Bank, 2013). There has been a significant increase in the number of people participating both directly and indirectly in the sector and this is expected to continue given the socio-economic realities of most developing countries. High levels of poverty and unemployment and growing inequality continue to be largest problems facing the majority of developing countries today. The ASM sector plays an important role as a source of livelihoods, particularly for those residing in rural areas with limited economic opportunities. In the majority of countries, the recognition of the ASM sector was driven by its increasing contribution to socio-economic development through job creation, poverty alleviation, and rural development. In response, a number of intervention strategies were introduced in support of the sector to address its negative impacts and to increase its potential benefits."

Jan 1, 2017

[Consider the situation where we wish to estimate the value of a block of ore W from the values of n samples wt (i = 1, 2.... n). The samples can be located anywhere inside or outside W. The blocks and the samples can have any size or shape (Fig. 8.1). We define ws as the set of all the samples wi: WS = Iw1; W2; W3.. . wn and we wish to estimate the block W using the samples ws. Let xi be the value of the ith sample and µw the unknown average value of the block. We can estimate µw from the average value of all the samples. Let µw be the estimator of µw. As the samples need not be of equal size, the sample average is: µw = wi xi/~ wi. (8.1) The mean squared error made when estimating µw by µw is known as the variance of estimation of W by ws. It is the error made when giving the value of the sample set ws to the block W and is written o2E (WS to W). By definition: o2E (ws to W) = E [(µw - µw)2]. (8.2) If n = 1, this error is known as the extension variance. Since the sum of the weights given to the sample values is equal to 1,0, the estimator µw is unbiased (the question of the absence of bias is studied in more detail in Chapter 9): E [( µw - µw)1 = 0. (8.3) The error (µw - µw) has zero mean and a variance o2E = E (W8 to W). If we assume that this error is normally distributed, we can define confidence limits for µw. Let tP be the value of the standardized normal variance t such that the probability that t is less than tP is p: (t,) = .\/27T f exp (- t2/2) dt = p (8.4) The probability that µw is larger than µw + t1-P oE is p. The probability that µw is smaller than µw- t1-P oE is p. The following approximation formulae are often used to calculate 2p central confidence limits for the block value: 68% central confidence limits (p = 16%): µw - o2E, and µw + o2E. 95 % central confidence limits (p = 2,5 %): µw - 2 oE, and µw + 2 oE. Tables of the function 0 can be found in any textbook on statistics (for example, Fraser, 1958, p. 388) (see Table 7.2). [ ] The last line of Table 2.2, corresponding to f = 00, also contains the value of tP for p = 80, 90, 95 and 97,5 %.]

Jan 1, 1978

By C. I. Ware

The phrase ?back to basics? is used to describe the fundamental principles and philosophies on which a successful mining operation is based. At Richards Bay Minerals (RBM), ?back-to-basics? (BTB) has been implemented as a continuous process to probe operational functions with a view to increase production without capital expenditure. It also ensures that the fundamentals of dredge mining are understood and adhered to. BTB is based on the principle that people drive organizations, necessitating an interactive implementation approach. At the RBM mine this entails a three-pronged approach, namely practical training, teamwork initiatives and the use of lean Six Sigma methodology. It is the latter on which this paper will focus. The lean Six Sigma methodology is successful in identifying, managing and documenting BTB projects, while team work initiatives are integral to successful project implementation. This paper describes two examples of how the Six Sigma methodology can be applied to a dredge mining operation. The first example focuses on the improvement of Mining Pond Echo (MPE) plant throughput by reducing the lost opportunity (non-production) due to tails downtime. The project was successful in increasing plant throughput by 4.5%. This was achieved by increasing the tails disposal system availability by running MPE as a four tails lines operation and focusing on causes of tails production downtime. Subsequently, the tails system at MPE is no longer the production bottleneck. Second, the Six Sigma methodology was successfully applied at Mining Pond Charlie (MPC) to increase the primary dredger throughput by 4.6% through focusing on the dredging technique and maintenance issues. This project highlighted the need for trend analyses to ensure an effective dredging technique, coupled with simple production documentation, which also ensures a focus on maintenance. Keywords: Back to basics, Six Sigma, training, teamwork, throughput improvement, tails, dredger.

Jan 1, 2007

By F. T. Cawood

This paper reviews the Mineral and Petroleum Resources Royalty Bill of 2006, identifies potential areas that require further consideration by the National Treasury and, where appropriate, offers alternatives. At its basic level a royalty regime consists of two elements: the base and the rate that must be applied to the base. Deciding on an acceptable royalty rate is a complex matter and can be done only after extensive analysis of all variables affecting the competitiveness of the regime. The proposed royalty rates are in the range of zero to five per cent and are based on sales revenue. It is argued that future investment patterns are more likely to be influenced by market cycles and commodity prices than policy events. Inevitably the current boom in the minerals sector will come to an end and the success of the proposed regime will be known only when profit margins are under pressure. Policy makers must therefore have the long run in mind when developing policies that affect the ability of industry to do business, without the need to renegotiate each time there is a change in the market. A long-run policy must be sufficiently flexible to allow for consistently fair sharing ratios and splitting of revenues between government and industry. Although broadly in line with international practice, the current version cannot be regarded as internationally competitive and efficient. In order for the royalty bill to become internationally competitive, promote beneficiation of mineral production and adequately compensate the State for the loss of a national resource, the National Treasury should heed the following main recommendations in this review: 1. Abandon the notion that each mineral product must have its own royalty rate; 2. Introduce an NSR type royalty that allows for value-addition costs to be deducted from the base; and 3. Provide for automatic relief through the use of a sliding-scale mechanism based on ability to pay, e.g. the X-Factor used in the gold mining tax formula.

Jan 1, 2007

It is almost eight years to the day when we launched the original SAMREC Code in March 2000. Since that date I think the SA minerals industry has matured beyond recognition in its approach to reserve and resource reporting, and I believe we can be justifiably proud of that progress. Many of those companies that have led the way in reporting standards are listed on our sponsors list. We thank them for their sponsorship but even more for their considerable support in all aspects of the SAMCODE organization. Matt Mullins, my co-speaker here today, has been instrumental in the revision of the SAMREC Code over a long period of time and also took up the reigns with the SAMVAL Code after the retirement of Rob Croll. I would like to formally thank Matt for his immense contribution in producing the two Codes; he can take great pride in the final product and fully deserves the many accolades that we have already received from our international counterparts. He has done all this whilst still maintaining a very demanding day job at BHPBilliton. Matt will be the first to admit that he has received tremendous support from more than 100 people who participated in the various working groups, many of whom are employed by our sponsors. There are obviously too many to mention by name but they deserve a big vote of thanks for their personal contributions. In addition to launching the two Codes we are also launching our ?branding? initiative with our own SAMCODE logo and the SAMCODE website: www.samcode.co.za. All the slides used in the presentation today will be on the website later today as will a newsletter which will serve as a media release.

Jan 1, 2008

By D. Rankin, H. J. Stucke, W. K. B. Loftus

SYNOPSIS Planning for the Finsch opencast mine is described with particular reference to the proposed stages of mining, the method used to determine the limits of surface waste rock stripping and the derivation of the waste rock mining rate. Actual mining practice is described and the possibility of installing a skipway into the open mine at a later date is discussed. In addition, the waste dumping policy is referred to: this includes the cutting of slots through the hill on the west side of the pipe to increase tramming speeds. The crushing, washing and recovery sections of the treatment plant are described together with the various changes being made to the circuit. The most important of these are the installation of X-ray sorters for the final recovery of + 7 mesh diamonds, and the vibrating grease belt for the recovery of - 7 mesh diamonds. INTRODUCTION The Finsch Mine, a subsidiary of De Beers Consolidated Mines, Limited, is situated approximately 90 miles west of Kimberley, and some two miles from the village of Lime Acres, in the Northern Cape. Lime Acres is a company village shared with the Northern Lime Company which operates a limestone quarry and works near the township. The mine is near the southern end of a well mineralized belt producing limestone, asbestos, diamonds, manganese and iron ore, and the whole area has developed rapidly in the last 20 years. Power for mining operations is drawn from the Electricity Supply Commission. The mine is conveniently close to the main Kimberley-Postmasburg railway line. Water is drawn from the Vaal River through a 57 mile State-owned pipeline at present temporarily operated by mine staff. DISCOVERY AND PROSPECTING OPERATIONS The diamondiferous orebody is situated on the farm Brits, which is State-owned ground that has been leased to the Company for the period of mining operations. There are indications of interest in the area by prospectors (Richter, de Bruin and others) over a period of some 30 years and the discovery would possibly have been made years earlier but for a law preventing prospecting for precious stones on State-owned land. Although the pipe formed a clear depression on the hill where it was situated, it was overlain by some 5 to 40 ft of banded ironstone rubble which was washed into the depression from the hillside above the pipe and obscured the deposit. In 1958, Mr Allister Thornton Fincham applied for, and was granted, the right to prospect for base minerals on the farm Brits. Asbestos is known to occur in the area and, initially, was Mr Fincham's primary interest. However, during the prospecting operations, garnets were discovered and it was realised that diamonds might be present. In 1960, when the Precious Stones Amendment Act was passed, a company

By V. L. Nkuna, S. R. Amos, T. Naidoo

"Ivanhoe Mines’ Kamoa copper project is a recently discovered high-grade copper sulphide deposit in the Katanga Province of the Democratic Republic of Congo. The mineralization identified to date within the resource is typical of sedimentary-hosted stratiform copper deposits and comprises three distinct units: supergene, mixed, and hypogene mineralization. Since 2010, various metallurgical test work campaigns have been conducted on drill core from different areas within the deposit in line with resource expansion. The most recent bench-scale flotation test work has shown positive repeatable results. The work was conducted on a composite sample of drill core from the southern part of the Kamoa resource. The sample is representative of the first four to five years of planned production, when Ivanhoe Mines intends to produce clean, highgrade copper flotation concentrate. A detailed test work roadmap developed from first principles for the campaign led to an optimum flow sheet for treating the Kamoa mineralization. Copper recoveries of 88.3%, at a concentrate grade of 39.0% copper, were achieved using a composite sample — an improvement on the previously published 85.9% life-of-mine average copper recovery indicated in the November 2013 Kamoa Preliminary Economic Assessment. IntroductionThe Kamoa project is a newly discovered, very large, stratiform copper deposit with adjacent prospective exploration areas within the Central African Copperbelt, approximately 25 km west of the town of Kolwezi in the Democratic Republic of the Congo (DRC) and about 270 kilometres west of the provincial capital of Lubumbashi. Ivanhoe Mines holds its 95% interest in the Kamoa copper project through a subsidiary company, Kamoa Copper SA Limited SPRL, and DRC government owns 5%. The project location is indicated in Figure 1.In January 2013, a new independent mineral resource estimate ranked Kamoa as Africa's largest high-grade copper discovery and the world's largest undeveloped highgrade copper discovery. At this time, Ivanhoe Mines had estimated discovered Indicated Mineral Resources of 739 Mt grading 2.67% copper, containing 43.5 billion pounds of copper, and Inferred Mineral Resources of 227 Mt grading 1.96% copper, containing 9.8 billion pounds of copper, as demonstrated in the Preliminary Economic Assessment (PEA) (AMC, 2013). A 1% copper cut-off grade and a minimum vertical mining thickness of 3 m were applied in each classification."

Jan 1, 2016

By A. Ndibalema

Blasting is one of the primary functions in any mining operation. Despite the fact that it generally constitutes between 30% to 40% of the mining cost, blasting also affects the cost-effectiveness of down steam activities such as load and haul, safety, tyre life for haul trucks, crushing and milling. Over the past three decades, significant progress has been made in the development of new technology in an attempt to reduce costs and increase efficiencies and productivities of blasting activities. Mine to mill project concepts have paved the way for such developments. Other advancements have included sophisticated computer modelling technologies for blast design and performance analysis. There are many factors that affect the cost of blasting. These can be divided into two categories; controllable and uncontrollable. The latter includes costs related to depreciation, exchange rates and government legislation with the former category including contractor versus owner blasting costs, efficiency and productivity issues and product quality. Controllable factors have one common denominator which can determine their outcome; direct site management influence. Excavator digging rates and final floor conditions give a good measure of blasting performance, and alternatively cost effectiveness. This paper is aimed at benchmarking against a blast optimisation database, understanding the relationship between costs and identifying opportunities for improvement. Understanding the cost drivers allowed a typical cost model to be developed for the current mining operation at Geita and this was used to drive continuous improvement with emphasis on closing the information loop.

Jan 1, 2008

By J. Grobler, J. Zietsman

"Understanding the influence of feed material mineral assemblage on spiral separation is paramount for the successful and consistent operation of a spiral plant. Plant stability is required to maintain high recovery of valuable heavy minerals (VHM). Standard material analysis methods used in the heavy mineral industry to evaluate spiral separation behaviour are unable to quantify performance differences in the case of large variations in feed assemblage. This study illustrates a systematic approach to evaluate spiral separation performance. This technique was applied on a four-stage spiral plant. The performance information was used in a circuit simulation to clearly illustrate the importance of correct setting of the spiral splitter position and resulting spiral mass loading to ensure optimal circuit stability and maximum VHM recovery.IntroductionHeavy mineral sands deposits generally consist of naturally fine-grained, free-flowing sand due to the way in which the deposit was formed. The size of the heavy minerals market is roughly 8 Mt/a of zircon and titanium combined (Iluka, 2013). If an average combined feed grade of 5% zircon and titanium and an average combined overall product recovery of 80% are assumed, the resulting run-of-mine feed tons that would require primary spiral beneficiation is 200 Mt/a. The number of spirals required for the processing of this material, assuming 6500 operating hours per annum, an average of 2 t/h per start, and the same number of rougher spirals as other duty spirals, is over 30 000 spirals for the heavy mineral sands business alone. These numbers clearly illustrate the importance of the spiral in the heavy mineral industry.The aim of a spiral plant in the primary processing value chain of a heavy mineral process is to reject non-valuable quartz (a light mineral) from various valuable heavy minerals (VHM). The product from the spiral plant is a heavy mineral concentrate that has a reasonable grade (more than 90% heavy minerals) at a reasonable heavy mineral recovery (greater than 90%). This benchmark is fairly easy to achieve if all the mineral particles are well liberated (as in the case free-flowing sand), the feed is properly deslimed (-45 µm fraction removed to less than 5% of the solids mass), and the spirals are fed within design solids concentration (water to solids ratio) and design spiral load (dry tons per hour per start)."

Jan 1, 2017

By R. E. Phillips, L. A. Cramer, R. T. Jones

Atomaer (Pty) Ltd is a technology-driven company with interests in Western Australian nickel. Atomaer (RSA) Pty Ltd is a process technology company operating in Southern Africa. It is a subsidiary of Atomaer Holdings Pty Limited (Atomaer), which has operating companies in Australia, Africa, South America, and Europe. In addition, Atomaer has substantial shareholdings in Braemore Resources Plc, a nickel company with interests in sulphide nickel projects in Western Australia and Southern Africa listed on the Alternative Investment Market of London Stock Exchange (AIM); Mintails Limited, a gold company with interests in gold operations in South Africa listed on the Australia Stock Exchange; and Brinkley Mining Plc, a uranium company with interests in Southern Africa listed on the AIM. Independence Platinum Limited (IPt) has been formed by Atomaer to undertake the commercial development and exploitation of Mintek's ConRoast technology with the strategic objective of providing independent smelting and refining facilities with lower capital and operating costs for the treatment and processing of platinum concentrates, including those with high chrome and/or nickel and/or low platinum group metal values, for the development of platinum projects by junior resource companies in South Africa. The platinum mines of South Africa have traditionally exploited the Merensky Reef of the Bushveld Complex for platinum production. However, the mining of this reef is reaching depths that require significant vertical shaft systems and costly refrigeration for access to the ore bodies at deeper levels. In the past decade, all of the platinum producers have begun to mine the UG2 reef because of its cheaper mining costs. UG2 ore usually has lower platinum group metal contents and higher chrome values. The new junior resource companies have gained access primarily to UG2 reefs. Consequently, platinum mining is becoming more UG2-based and the resulting concentrates contain high levels of chromitite. UG2 concentrates are difficult to smelt in the current six-in-line submerged arc furnaces, and this leads to increased capital and operating costs for those smelting operations. The platinum mines of South Africa produce the majority of the world's platinum metal used in automotive catalysts and jewellery. As the platinum market is forecast to grow at a healthy rate into the future, the SA producers Anglo Platinum, Impala, Lonmin, and Northam are planning to expand mining operations in the next decade. Furthermore, a number of smaller new producers who are accessing other mineral resources are entering the buoyant market and will produce platinum concentrates for purchase or toll treatment. Hence, a shortage in both smelting and refining capacities is anticipated. Further smelting and refining facilities independent of the major platinum producers will be required in South Africa by 2010. Mintek's ConRoast technology is best suited to fill the need to treat a high proportion of UG2 concentrates in a cost-effective and technically efficient manner. Atomaer base metals technology has been developed for rapid leaching of nickel and other base metals from alloy. Significant reductions in capital and operating costs are expected from the combination of the two advances in process technology compared with that installed at the existing smelters and refineries that were established over 20 years ago. Independence Platinum will be listed on AIM and the Johannesburg Stock Exchange (JSE) in 2007, and developed as an independent, vertically-integrated platinum company using its independent smelting and base metals refining facilities as a strategic platform to promote the development and mining of platinum projects by junior resource companies in South Africa through technical co-operation, strategic alliances, joint ventures, off-take, toll-refining, and marketing agreements.

Jan 1, 2006

By R. Tabanrad

This article examines the stability analysis of intake tunnels of Karun III dam & power plant. The tunnels have been excavated with a 3-dimensional and fairly complicated geometry in Pabdeh weak formation rocks at Zagros region. Due to the particular excavation geometry and inapplicability of plane-strain status, the stability of these excavations has been analyzed using both the 3D elastic software Examine 3D and the 2D non-elastic software UDEC simultaneously. The geomechanical parameters have been estimated by back analyzing the measured displacements from instruments, which were installed in the tunnels, and the parameters have been used in non-elastic numerical model. The results of the analysis show that the strains occurred in surrounding rock mass are less than the allowable failure strain and that the tunnels are stable. In addition, considering the results of non-elastic 2D analysis, the areas with unstable potentiality were identified and after the instability occurred at the areas predicted by the model , it was stopped up to the allowable strain limit through installing additional support . Bei diesem Artikel handelt es sich um eine Analyse der Festigkeit der Eingangstunnel des Damms und Elektrizitätswerks “Karoon III”.Die Tunnel haben drei Dimensionen und eine sehr schwierige Form und sind in brüchigen Felsen in “Pabedeh ,in Zagrossgebirge”eingebohrt.Aufgrund besonderer Form der Bohrung wurde ihre Festigkeit durch elastische dreidimensionale Soft-ware analysiert,die gleichzeitig die drei-und zweidimensionale unelastische Soft-ware UDEC benutzt hat.Die geometrische Parameter sind mit Ausnützung der Analyse berechneter eingesetzter Versorgungen,die in Tunnel aufgebaut sind,bewertet.Und diese Parameter sind in Zahlen-Analyse benutzt.Die Ergebnisse dieser Analyse zeigen,dass die Drücke an Felsen weniger als problematischer Grad der Drücke sind und die Tunnel fest sind.Ausserdem wurde mit einem Einblick auf die Ergebnisse der unelastischen zweidimensionale Analyse die eventuellen unfesten Teile bestimmt.Nach entstandener Unfestigkeit an diesen Teilen wurde die problematischeDrücke für Aufbau mehrere Anlagen gestoppt. Cet article examine l'analyse de la stabilité de tunnels de la prise de Karoun barrage III et centrale électrique. Les tunnels ont été excavés avec une 3 dimensionnel et compliqué géométrie dans Pabedeh que la formation faible balance à région Zagros équitablement. Dû à la géométrie de l'excavation particulière et inapplicabilite de statut de tension plane, la stabilité de ces excavations a été analysée en utilisant les deux le 3D logiciel élastique Examen 3D et le 2D logiciel non - élastique UDEC simultanément. Les paramètres du geomechanical ont été estimés en analysant le déplacement mesuré d'instruments qui ont été installés dans les tunnels et les paramètres en arrière a été utilisé dans les modèles numériques non - élastique. Les résultats de l'analyse montrent que les tensions se sont produites dans masse du roc environnante est plus petit que la tension de l'échec admissible et que les tunnels sont stables. De plus, les régions avec potentialité instable ont été identifiées étant donné les résultats de 2D analyse non - élastique, et après que l'instabilité se soit produite aux régions prédites par le modèle, il a été arrêté jusqu'à la limite de la tension admissible à travers installer du support supplémentaire.

Jan 1, 2003

By D. Borman

The location of underground positions in mines has usually fallen to their survey department's. Their usual tasks include: ? installation of a high-precision peg network like the National Survey grid as beacons ? offset measurements to the sidewalls, and sometimes hangingwall and footwalls, of development tunnels and excavations ? offsetting of stoping panels for the calculation of face ? advance and the square meterage of area removed. Other departments build their observations into the plans produced by the surveyors. If more precision is required, they ask the surveyors to `elevate' their points with a theodolite. Apart from the survey peg network and `elevations' these measurements are usually planer measurements suitable for producing plans, sections and projection sheets. When the only representation medium was a sheet of paper or transparency, this was perfectly suitable. However the development of CAD models of the workings has changed all this. Full 3D representation is not properly catered for in the tool sets in current use, and needless hours are spent converting analogue-type observations into digital form. The basic survey models need to be augmented by the observations of other disciplines that are also required to locate positions underground. These include: ? the need for geologists to plott their mapping and ? borehole logs into 3D space ? the need for samplers to position their sample points accurately because the geostatistical models are biased and inaccurate ? the need for stoping width recorders to position their thickness observations to prevent errant evaluation and bonuses ? the need for rock engineers to place their mappings into 3D to build true geotechnical models to manage geotechnical stresses ? the need for production officials to receive updated face ? positions during the month so that they can manage the ? production and utilize their resources properly. This paper covers a new method of positioning profiling points to a suitable degree of accuracy. I believe the method is sufficiently simple and cheap for all to use. Historical methods Platinum and gold orebodies, especially in South Africa, are generally narrow tabular deposits. This made them suitable for depicting on a 2D medium like paper. The determination of stoping or face widths was not a problem, as the ore is removed in a single stoping process. Such 2D representations could be easily positioned using tape offsetting between survey pegs or points tied in from survey pegs. Problems of dip were handled by using 'Stope Sheets' which chose a best fit projection for the whole area depicted or by using true dip section sheets for inclined tunnels. Rolling reef and potholes were the common areas of difficulty. Most mines now model their orebody and plan the extraction in computer graphics. This means that the representation of the current workings has to change to support this trend. It now becomes a truly 3D problem. Mining officials find it inconvenient to have their actual workings in single-line depictions while their orebody is depicted with top and bottom contacts and the planned tunnels in 3D shapes. Geological models which form the basis for the resource model and planning process are often out of date, as the updating process is slow. The geological mapping has to be plotted onto sections and plans. These are then digitized into graphics by a draughtsman and checked by the geologist. Only then does the interpretation and model update take place. Potential methods to create 3D 'Actuals' Global positioning systems (GPS) Surface mines have been blessed by the advent of GPS technology. These simple-to-operate pieces of equipment and compatible software make it possible for all types of mine employees to locate themselves, and therefore to locate their observations. GPS are often inexpensive and give an acceptable degree of accuracy. Only in deep pits or close to the highwalls are observations hindered. Such technology cannot however, work underground. Ground penetrating radar (GPR) GPR is reported to be very useful for estimating the positions of potholes and faults. However it is still considered as providing an estimate, not a definite measurement. GPR is good for seeing into solid rock, not into open spaces. Inertia reading devices Experiments with inertia reading devices have been conducted on the mines. They were the forerunners of GPS for military guidance systems, and were even used for city navigation in the early 70s. They do however work best with heavy or fast-spinning gyros. Anglogold Ashanti's attempt to use them were discontinued in the mid-90s when the developers asked for an additional few million rand for a device that would weigh more than 12 kg and have a limited operational life.

Jan 1, 2006

By M. J. Deats

INTRODUCTION To prepare the reader adequately, reference should be made to the paper 'The Pioneering of Fully Mechanized Longwall Coal Mining in South Africa' by R. T. Naude and M. J. Deats, published in the February, 1967, edition of the Journal. At the time mechanized longwalling had been proved operationally feasible with encouraging results since its inception in South Africa in mid 1965. Subsequently, further top seam panels were successfully extracted with improved productivity and with some modification to equipment. In 1968, after an overseas visit by mine officials, equipment specific to suit local conditions was acquired and lower seam trials commenced. Despite initial obstacles and difficulties the first panel was successfully mined. Modification and improvements to equipment and techniques resulted in the second lower seam panel being extracted economically during 1969-1970 and with productivity exceeding even that of top seam faces. During extraction of the second lower seam panel, gradual and complete surface subsidence occurred over the panel without affecting face operations. Rock mechanics investigations had originally indicated some uncertainty regarding adverse dolerite sill behaviour which could have caused longwall mining to be hazardous. All such doubts have now been removed. RESUMÉ AS AT FEBRUARY, 1967 Mechanized longwalling was introduced in 1965 on an experimental basis in an effort to improve productivity in the newer area of the mine and to increase the quality of the comparatively dirty run of mine feed to the washing plant. In addition advantages could be predicted in the long term by better utilizing reserves of straight coking coal and increasing the life of the mine. The top seam unit was obtained on a rental basis and by the beginning of 1967, two panels, the first having a face length of 215 m, a panel length of 345 m, seam section of 1 219 mm, and the second the same face length, a panel length of 453 m and a seam section of 1 067 mm had been successfully longwalled. At this stage a best month of 24 602 metric tons with an average of 17 364 metric tons per full working month could be reported. Results were sufficiently encouraging for the company to exercise its right to acquire the equipment as its own asset and from 1967 onward, the unit has operated on this basis. From the point of view of rock mechanics, the dolerite sill over the first two panels was known to have 'bridged', only the material below the base of the sill having truly 'goafed'. No adverse pressure effects had significantly affected face operations except for two 'bumps' towards the end of operations in each panel. Gate road maintenance presented no problem. The strong inflows of water experienced on the first face were now known to be associated with an isolated water bearing fault plane and were not experienced on the second face. TOP SEAM EXPERIENCE - MAY, 1966, TO DATE Panel 2 (February, 1966, to November, 1966) The operations in panel 2 (see Fig. 1 for location of panels) continued uneventfully until the final month in this panel. At this stage, the chocks behaved most inconsistently often lowering under the weight of the canopies alone. Roof trouble became so severe that the unit was unable to produce adequately and blasting of chocks became a daily occurrence. Finally 24 m short of the planned limit it was decided that a complete overhaul was the only solution and the extraction operations commenced. At this time, the only fatal accident associated with longwalling occurred when a Non-White inadvertently fell into the panzer conveyor transfer point at the main gate. Some gate road difficulties in this panel were overcome by bolting tapes to the roof as decking and allowing them to pass over the gate chocks and collapse into the goaf. Two cribbed roof falls were also negotiated by the gate chocks at the face ends without undue difficulties. It was decided in the light of maintenance problems encountered on this face that the best solution would be to appoint a foreman solely in charge of face maintenance crews. This step has to a large extent contributed to success achieved subsequently. Panel 2-face length 213 m panel length 456 m cutting height 1 092 mm average monthly output 16 443 t best month 19 130 t Panel 3 (December, 1966, to June, 1967) By the end of 1966 results were sufficiently encouraging for the company to exercise its right to acquire the equipment. This was negotiated and the unit has operated as such to the present date. Prior to installation in panel 3, all face equipment was thoroughly overhauled underground and the chocks boost-tested to yield-load. The costs of overhauling, including almost complete hose replacement, were much higher than anticipated. Advice had been received that complete hose replacement was common practice after +/- 18 months' operation. Shortly after commencement of operations in panel 3 the chocks again failed as they had towards the end of panel 2. At this stage, the suppliers advised that the valve gear should have been repaired during the overhaul as the needle and seat seals could no longer be relied upon. This operation was impossible to carry out on the face and would have been e1ttremely time consuming. Fortunately a new type of valve gear was then available from the suppliers, termed the capsule-type valve, and a complete change over could be made on the face. The encapsulated valve has a delrin pad sealing over a raised

By R. R. M. Cousens, W. S. Garrett

Discussion: L. T. Campbell Pitt: When Mr Garrett and I discussed the design and construction of plugs and bulk head doors to resist great water pressures in mines about ten years ago, we visualized an emeregeny when there would be no time for site preparation. We could not however, anticipate a problem of such magnitude and urgency as the one described in the paper under discussion. The paper very clearly describes the flooding and the measures taken to save the mine. The complete flooding would have been a national disaster. The account of the combat between man and water reads like the account of a military operation. It was an endeavour which achieved success because there was a master plan, tenacity and, in fact, heroism. The paper is of necessity a long one. In my contribution I will mention some historical facts and factors that I believe have relevance to the flooding and the saving of the mine. It is of great satisfaction to me to have been one of the two initiators of a new approach to underground plug design and construction. The investigation was taken over by a Chamber of Mines Committee which combined our early work with further research and the resulting data had a significant bearing on the West Drie emergency plug designs. The final plug length was correctly based on the requirements to obviate leakage. A plug to have structural strength only would have been very much shorter but so ineffective as to have become a waste of labour, material, and in consequence, wasted the limited time available to prevent flooding of the whole mine. The figure of 75 lb per sq in. sheer stress provided a very ample margin, in my opinion, but under the very difficult site conditions and the limitations of plug material suitable for the very unusual construction, justified such a figure. There must have been considerable temptation to reduce the safety margin due to the urgency to complete and commission the plugs, but it is much to the credit of the directors of this operation that nothing less than a safe design was permitted. In about 1945 Blyvooruitzicht Gold Mining Company, which was then mining in to the West Driefontein area to assist in early development, decided to protect themselves from a possible inrush of water from the latter mine. A bulk head door was accordingly designed and installed on 2 level between 1945 and 1948 at 4,218 ft below surface. Since West Driefontein paid for it, Gold Fields consulting mechanical engineers collaborated with their colleagues of Rand Mines who completed their design after a few modifications were agreed. This door was never used to prevent an inrush into Blyvooruitzicht but the exercise was very useful to Gold Fields when the high pressure doors initially installed on 10 and 12 levels between Nos. 2 and 3 shafts to enable isolation of one section of the mine from another were designed. At that time the capacity of West Driefontein to cope with an inrush of water was comparatively small because the storage in stoped areas and the installed pump capacity were much less than in later years. These doors together with others of the same design were later moved to the lower levels served by sub vertical shafts. In addition, doors to prevent flood water entering the pump chambers on 18 level were installed at each end of that level, and designed for a head of 360 feet so that water would have to reach 16 level before the pumps were flooded. All these doors were called upon to hold water and it is reported that only a small plug in one of the doors and a 4 in. dia flanged pipe joint failed. The lesson is that in these cases the smallest details warrant as much attention as the major ones. That these were the only equipment failures when so much depended on equipment, is very much to the credit of design engineers, both of the Gold Fields head office and companies who supplied both mechanical and electrical equipment. Until I had read the paper I wondered why the policy of isolation by bulk head doors was not continued as the development moved eastward towards the Bank Compartment and particularly when it pierced the Bank dyke. The reasons have been given in the paper. It was unreasonable to expect provisions to be made to counter the remote contingency that the large installed pumping capacity and mine capacity to store water would be exceeded during a peak inrush. When pumps were being considered for West Driefontein, it was decided that the Gold Fields standard underground main pumps at that time of 1 million gallons per 24 hours would be too small. Three alternatives were considered. 1. Very large units that would require underground sub assembly and dismantling of both pump and motor. 2. The largest pump and motor that could be conveyed as a complete pump or motor from surface to the pump chamber. 3. Pumps of higher speed to reduce dimensions and weight. It was decided to adopt alternate 2 because of avoidance of transporting pumps in parts and assembling underground. Alternative 3 was not favoured because high speeds would have introduced problems of more exact rotor balancing. I believe the decision was correct and in the crisis enabled comparative easy transport of pumps and motors to surface for overhaul and, if required, transfer to other pump stations. The standard pump thus became a 10 stage 1,500 rpm pump with a capacity of approximately 2 million gallons per 24 hours against a static head that varied

Jan 3, 1969

By C G. Robinson, B. K. Loveday, J. C. Paynter

"A method of optimizing the design of a processing plant has been developed which minimizes capital and operating costs for average stripping ratios and ore characteristics. Models were developed to describe the operating characteristics and costs involved in various operations such as up-grading, leaching and solid-liquid separation.The options available in up-grading policy and the relation between the cut-off grade and the costs incurred subsequently in the treatment of the ore formed the basis of the optimization. To arrive at an optimum plant configuration and a set of operating parameters, the interrelation between grinding size, leaching temperature, leaching time, and various plant recycle options was investigated."

Jan 1, 2014

By N. Hanief

This research involves the investigation of phase transformations in the platinum-chromium coated system. Single-layer 0.1 µm platinum coatings were deposited via electron beam deposition on 99.98 percent pure chromium substrates. Specimens were subjected to systematic heat treatment in a vacuum furnace at 900°C for up to 20 hours. Phase formation and the changes in surface morphology were investigated by X-ray diffraction (XRD) and scanning electron microscopy (SEM).Both CrPt and Cr3Pt phases are formed during heat treatment for different times at 900°C. Significant changes in the morphology of this coated system were detected after heat treatment at 900°C for 20 hours.

Jan 1, 2012

By Daniel W. Kappes, Pierre F. X. Mousset-Jones

"In this paper a discussion is presented of a program designed to produce a 40-day production forecast for an underground stoping system.The program is intended to be a tool for the operations department in an underground mine, and is designed to utilize data readily available from production reports. Brevity of data input, ability to revise slope conditions easily and the accuracy of starting stope status are stressed. Applicability of the program to a wide variety of sloping methods makes the program potentially useful in a large number of mines.The general functions of the program subroutines, and input and output lists are discussed. An example is also given of a stope in operation for a three-day period, which shows how the program can handle production irregularities as they arise."

Jan 1, 2014

By Z Bagheri, E Moosavi

Identifying the exact borders of geological domains and assessing the uncertainty of these borders are crucial steps in resource estimation. This potential field method was designed to build 3D geological models from data available in geology and mineral exploration, namely the geological map and a digital terrain model, structural data, borehole data, and the geologist’s interpretations. The resource evaluation of a copper ore deposit is often performed in three steps: (i) delimitation of the boundaries of the units corresponding to the various geological formations or ore types; (ii) estimating densities; and (iii) estimation of grades within each unit. In simple cases (e.g., a series of sub-horizontal layers), the geometric model can be built using 2D geostatistical techniques (kriging or co-kriging of the elevations or thicknesses of the various horizons) which also quantify the uncertainty of the model. A 3D model of Cu grade is generated using ordinary kriging; however, the indicator kriging method is run to design a 3D model of rock types by incorporating the drilling results. Block models of geophysical and geological characteristics are included in a similar 3D mesh to investigate their relationships in copper mineralisation.

Sep 1, 2021

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

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

Jan 1, 2020

By S. Kok, D. Delport, J. Calitz

Altering the microstructure in order to improve the tensile properties of bow shackles resulted in inconsistency in the fatigue performance. This raises the question whether the inconsistency in fatigue life can be attributed to microstructural changes along the profile of the shackle or to decarburization at the surface. Bow shackles forged from 080M40 (EN8) material were subjected to different heat treatments in order to alter the microstructure. The shackles were subjected to five different fatigue load cases, which represented typical loads experienced at termination points for an overhead power line with a span length of 400 m, with changes in conductor type, configuration, wind, and ice loading. Although the change in microstructure does improve both the tensile and fatigue performance, we found that the depth of the decarburization layer has a greater effect on the high cycle fatigue life of bow shackles than the non-homogeneous microstructure.

Oct 1, 2021