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By M. C. Bekker, T. Govindasamy

The potential benefits of implementing the Last Planner System (LPS) on mining infrastructure projects were explored through two case studies in the South African mining industry. The first part of the study involved measurement of the impact of LPS through earned value techniques on secondary project data. The second part sought to establish the lean construction success factors leading to improvements in construction performance. This was done by means of a research questionnaire distributed to the project owner’s teams, contracted parties, and the LPS facilitators. The results revealed a positive correlation between LPS application and planned percentage complete. However, the performance achieved during the LPS pilot was not sustained. The success factors and benefits considered most evident in the case study pilot differed among the three stakeholder groups. The findings are expected to guide construction stakeholders to better define performance measures and focus on factors required to make LPS implementation more effective in the South African mining sector.

Feb 12, 2024

By DeVerle P. Harris, David E. Euresty

"Potential metal supply refers to metals to be produced from as yet unknown deposits. Naturally, to measure potential supply requires, among other things, inference to the number of deposits, their tonnage, and their grade Thus, probability distributions were constructed for each 400-square mile cell in Northern Sonora, Mexico. A computer system was developed that simulated the occurrence of deposits for each cell, exploration for the deposits, and their exploitation. Evaluation was performed for each deposit by discounted cash flow analysis.An objective of this study was to explore the effect of the transportation network and required road or railroad construction upon potential supply. Consequently, the evaluation system was designed to consider the spatial location of a simulated deposit to the existing transportation network and the market destination. The construction of linkages to connect the deposit to existing transportation arteries as well as the carrying of the concentrates to the intended market were simulated. in terms of cost. The selection of the path took into account the existing terrain. This was done by digitizing topographical maps and representing the topography of the area by an orthogonal grid of elevations. The optimum route was selected by dynamic programming, meeting a prescribed gradient constraint. The potential supply of each cell was determined, allowing for the infrastructure effect."

Jan 1, 2014

By R. Koen, H. Visser, A. Roodt

"SynopsisThe tetraethylammonium salt of the mono-anionic coordination compound mer-oxidotrichlorido(thenoyltrifluoroacetylacetonato- K2O,O’)niobate(V) (NEt4)[NbOCl3](ttfa)], has been prepared under aerobic conditions and characterized by single-crystal X-ray diffraction. (NEt4)[NbOCl3](ttfa)] crystallized in the monoclinic P21/c space group, with a = 11.483 (5), b = 12.563 (5), c = 17.110(5) Å, and ß = 100.838 (5)o. The complex structure exists in a 50.0% (NbA) : 50.0% (NbB) positional disorder ratio.IntroductionComplexes containing organometallic type ß- diketone ligands with O,O and O,N donor atoms are used widely in coordination chemistry and have applications in catalysis, radiopharmaceuticals, etc. (Schutte et al., 2011; Roodt, Visser and Brink, 2011; Brink et al., 2010; Otto et al., 1998). These ligand systems are very useful because of their highly coordinative nature, high solubility, and also due to their ability to be functionalized with various substituents on the carbonyl carbon atoms (Schutte et al., 2011).Only a small number of ß-diketonate ligands have successfully been coordinated to a Nb(V) metal centre, with only a select few being characterized by X-ray crystallography (Viljoen, 2009; Bullen, Mason and Pauling, 1965; Preuss, Lamding and Mueller-Becker, 1994; Funk, 1934; Davies, Leedlam and Jones, 1999; Allen, 2002. The synthesis and crystal structure determination at room temperature of mer-oxidotrichlorido (thenoyltrifluoroacetylacetonato- k2O,O’)niobate(V) {(NEt4)[NbOCl3(ttfa)]} was first reported by Daran et al. in 1979. Accordingly, for this current investigation, (NEt4)[NbOCl3(ttfa)] was re-evaluated at 100 K to determine if the structural features might change with temperature."

Jan 1, 2015

By K. Isaac, S. O. Bada

This research focuses on the co-firing of discard coal with refuse-derived fuel (RDF) to utilize this abundant resource in South Africa for energy generation and reduce the volume of waste disposed of at landfills. The potential of a coal with a high ash content (> 40%), which is a grade used in some power stations in South Africa, and its combustion compatibility with two different RDFs in terms of emission reductions has been established. Gaseous emissions and ash residues from the combustion and cocombustion of the coal, two different RDFs and coal/RDF blends of different proportion were analysed. One of the RDF samples contained mostly paper (PB) and the other mostly plastic (PL). Co-combustion ash from the discard coal and RDFs showed a decrease in chloride and alkali metal contents as the coal ratio in the blend increased. The slagging propensity of the co-fired blends was found to be very low, while the propensity for fouling decreased from high to medium for all the blends with < 75% RDF. Cocombustion of RDF with coal showed a decrease in SO2 emissions from 387 ppm (discard coal) to 50 ppm for the sample containing 25% coal discard plus 75% PL. A 15% PL to 85% coal blend also reduced NOx emissions from 145 ppm (100% PL) to 88 ppm. The lowest CO2 emission observed was 6000 ppm for the blend of 85% discard coal plus 15% PB. It was established that the most favourable fuel blend that can produce the lowest sulphur emissions if used for power generation is the 25% coal discard plus 75% PL sample.

Aug 2, 2022

By T. Mahlangu Mahlangu, T. Fungene, K. C. Sole

"Although oxygen is widely employed as the oxidant of choice in gold leaching by cyanide, its low aqueous solubility presents some drawbacks in practical application; hydrogen peroxide has therefore been considered as a possible alternative. The aim of this investigation was to study the catalytic decomposition of hydrogen peroxide, which generates an oxidizing intermediate species, and to understand its effect on cyanide destruction. Operating conditions that facilitated the effective decomposition of hydrogen peroxide were established by varying the pH and catalyst type and concentration. The oxidizing intermediate, detected using an indirect technique, was found to be the hydroxyl radical (OH?). OH? is commonly generated in acidic solutions, but this work demonstrated that it is also produced at the alkaline pH values necessary for cyanide gold leaching. The effects of free and complexed iron and copper catalysts on the oxidation and consumption of hydrogen peroxide and cyanide were also investigated. It was shown that the cyano complexes of Fe(II) and Cu(I) are also effective as decomposition catalysts. Hydrogen peroxide concentrations above 0.01 M decreased the free cyanide concentration, which was attributed to the probable formation of the cyanate anion (CNO-). Although cyanide consumption increased due to its oxidation in the presence of OH?, excessive cyanide consumption in the presence of copper was attributed primarily to its complexation by the unstable copper(I) cyanide species. Rate constants for the decompositions of H2O2 and cyanide by ferrocyanide and copper cyanide were calculated; the latter was identified as being a better catalyst. IntroductionLeaching of gold from its ores using cyanide has been employed for well over a century (Habashi, 2016). Addition of cyanide converts the gold into a cyanide complex (Au(CN)2 -) that is soluble in water (Wang and Forssberg, 1990). The leaching reaction can be represented as Equation [1]:"

Dec 1, 2018

By J. K. Pothal, S. Basu, R. Barik, P. K. Senapati, A. Dubey

The aim of this paper is to study the characteristics of limestone particles (size varies from 4 mm to 20 mm) during its transportation from open cast mines through pipes. The influence of four different size fractions of limestone samples (4–6 mm, 6–10 mm, 10–15 mm, and 15–20 mm) on pipe flow characteristics during vertical transport in a 100 mm NB hose pipe was investigated. The flow velocities in the vertically lifting pipe was varied from 0.95 m/s to 2.78 m/s and a rotary feeder of 2 to 4 tonne/hr. capacity was employed to feed the limestone samples to the lifting pipe. The pressure drop data for the different size fractions of limestone samples at different mixture velocities were analysed. The effect of input volume fraction and superficial velocity of limestone samples on hold-up ratio were discussed. It was indicated that there was no substantial increase in pressure drop of slurry samples in the studied range of concentrations (CW = 4.22% - 12.54%), particle size, and low to medium range mixture velocities with respect to water run head loss data. The hold-up for the larger size fractions (15–20 mm) of limestone samples was found to be significant as compared to smaller size fractions, which may be due to terminal settling velocities of individual fractions of the studied samples.

Mar 4, 2025

By F. P. Janse van Rensburg, P. J. Le Roux

"There are numerous factors that affect open stope stability and often result in falls of ground. These falls of ground can be attributed to factors such as beam failure due to a larger than normal roof area (hydraulic radius too large), adverse ground conditions, seismicity, the stress-strain environment, absence of support, and poor drill-and-blast practices. The effect of time on the stability of open stopes is sometimes underestimated and is relatively unknown, especially on Target Mine. Actual data was collected from open stopes at Target and analysed to show the effect of time on open stope stability. The benefits of this analysis will include improved understanding of time-dependent failure, which can assist in reducing dilution and the risk of sterilization of future mining blocks. IntroductionTarget Mine is situated adjacent to the town of Allanridge, some 20 km from Welkom in South Africa’s Free State Province, and is the most northerly mine in the Welkom Goldfields area (Figure 1). The mine consists of a single surface shaft system with a sub-shaft (Target 1C shaft) and a decline. Ownership was attained by Harmony Gold Mining Company Limited in May 2004 (Harmony Gold Mining Company Limited, 2010). Before discussing the selection of open stopes for the back-analyses, a brief explanation of open stoping as practiced at Target Mine will be given. The orebody is some 180 m in thickness and 270 m wide and comprises multiple reefs overlying one another. The 180 m thick reef package is termed the Eldorado Reefs, as shown in Figure 2. The Eldorado Reefs suboutcrop against the Dreyerskuil Reefs, as shown in Figure 3. The dip of the reef varies from as low as 10° in the west to 75° in the east.Compared with most Australian and Canadian open stoping mining operations, Target Mine is unique. In most Australian and Canadian mining operations the hangingwall and footwall of the open stopes comprise waste rock, and the orebody dips relatively steeply. Due to the depth of Target Mine, some 2300 m to 2500 m below surface, a destress slot (Figure 4) is mined to create an artificial, shallow mining environment whereby the field stress is managed and kept at values of around 60 MPa. The destress slot comprises a narrow tabular stope with an average width of 1.5 m, and is mined on the Dreyerskuil Reefs.At Target Mine the hangingwall, sidewalls, and footwall of the open stopes all comprise reef of different grades, except for the EA1, where the EB footwall is waste rock. If a stope, for example, is mined alongside an existing old stope, the western sidewall of this stope will consist of backfill. The general mining direction of open stopes is from the lowest position of the reef on the west, progressing up towards the east as shown in Figure 5."

Apr 1, 2019

By S. C. Maji, U. S. Chattopadyay, S. Chaudhuri, T. Gouri Charan

India is devoid of coking coal resources, which is a vital raw material for coke making. As a result it has to import about 80% coking coal for steel making. Indian National Steel Policy, 2017 has envisaged crude steel production around 300 million tonnes Mt) by the year 2030-31 and forecasts coking coal demand of 161 Mtpa (million tonnes per annum). One way of improving the indigenous resources is to study the performance of existing washeries and to improve the clean coal recovery by adopting suitable technology. A study was carried out, mainly to focus on the fine coal circuit of an existing coking coal washery, to generate the desired quality of clean coal. The operating coking coal washery being studied is presently feeding coal from nine different sources with varying feed ash contents. The coal fines varied in ash content ranging from 19.5% - 43.6%. For treating the coal fines, the washery has operated a spiral for the past few years. Performance studies were carried out, and it was found that the spiral is not efficient in recovering the cleans. Detailed investigations were carried out on the coal fines using flotation techniques, and it was found that to achieve a low ash clean coal, the froth flotation is a much better beneficiation option than a spiral, The studies showed that the recovery of clean coal through a spiral is 67.7% at ash content of 27.0%, while flotation tests showed a recovery of 68.1% at 16.5% ash content. Keywords: Coking coal washery, near gravity material, spiral washer, froth flotation

Jan 1, 2020

By J. L. Gaigher

A simple mineralogical regression model is described, which was devised in order to help the flotation plant cope with rapidly changing headgrades and feed mineralogy. Minerals in the feed were determined by means of quantitative X-ray diffraction and compared with plant performance parameters and flotation reagent consumption rates, using regression analysis. A mineralogical regression model for predicting recovery served to motivate flotation plant personnel and also set standards whereby flotation performance could be measured. The reagent consumption model indicated in which direction any or all of five flotation consumption rates should be altered. All practically measurable parameters, even those showing poor or no correlation, were included in the model, provision being made for the continuous update of a batch of data of 30 lines. Compared with the method of multiple linear regressions, the r2 is poorer at 0,41 (versus 0,50), but the mean standard error of estimate for both methods is identical at ± 6,6 units. The regression model is simpler than the method of multiple linear regressions, and although the accuracy of prediction is not high (41%), the mineralogical regression model proved to be successful in practice. Flotation efficiency, as measured by the mean ratio of head to tails grade, for the 9 month period after the mineralogical regression model was implemented on an hourly basis, improved from 2,8 to 3,2, which is equivalent to an improvement in average flotation recovery of 5% relative.

Jan 1, 1987

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 N. T. Beukes

An engineering house?s perspective of required inputs in designing a copper electrowinning tank house and ancillary equipment calls for both understanding of the key fundamental controlling mechanisms and the practical requirements to optimize cost, schedule and product quality. For direct or post solvent extraction copper electrowinning design, key theoretical considerations include current density and efficiency, electrolyte ion concentrations, cell voltages and electrode over potentials, physical cell dimensions, cell flow rates and electrode face velocities, and electrolyte temperature. Practical considerations for optimal project goals are location of plant, layout of tank house and ancillary equipment, elevations, type of cell furniture, required cathode quality, number and type of cells, material of construction of cells, structure and interconnecting equipment, production cycles, anode and cathode material of construction and dimensions, cathode stripping philosophy, plating aids, acid mist management, piping layouts, standard electrical equipment sizes, electrolyte filtration, impurity concentrations, bus bar and rectifier/transformer design, electrical isolation protection, crane management, sampling and quality control management, staffing skills and client expectations. All of the above are required to produce an engineered product that can be designed easily, constructed quickly and operated with flexibility.

Jan 1, 2009

By X. Zhang, Z. T. Yuan, C. Zhang, H. M. Zheng, X. M. Wang, L. X. Li, N. Li

In metal mines, comminution is commonly used to liberate minerals for subsequent mineral concentration processes, such as flotation, magnetic separation and gravity separation. However, ores are aggregates of valuable minerals and their disseminated gangue minerals, and these minerals exhibit diverse comminution behaviours. In carbonate-bearing iron ores, hematite, siderite, and quartz are the predominant minerals, of which hematite is the objective recoverable mineral. In this regard, the present study aimed to identify the imperfections in siderite minerals by X-ray computed tomography (XCT), and thus quantitatively characterise their size and shape in terms of Feret diameter and aspect ratio, as well as other parameters. XCT images indicated the protogenetic ubiquity of pores <5 μm in diameter and cracks of 7-8 μm inside siderite particles. In the ground product of artificially mixed siderite and hematite, siderite particles were smaller than hematite particles. Siderite particles had a larger surface area and greater surface roughness than hematite particles of the same volume, which was attributable to the intrinsic holes and cracks in siderite particles. This study provides an insight into the relationship between the inherent imperfections of minerals and their comminution properties. Keywords: Siderite; X-ray computed tomography; particle imperfections; grinding behaviour; sliming

Jan 1, 2020

By I. J. Kotzé, R. T. Jones

Conventional PGM matte smelting requires the presence of a certain quantity of base metal sulphides in order to collect the platinum group metals in a molten sulphidic matte phase in the furnace. Furthermore, the quantity of chromium oxide in the feed materials is controlled in order to avoid the build-up of high-melting chromite spinels in the furnace. Mintek has developed a process for smelting difficult PGM-containing feed materials that contain low amounts of sulphur and/or high amounts of chromium oxide. A DC arc furnace is used to provide the appropriate conditions to generate an iron alloy that collects PGMs very effectively, leaving extremely low residual quantities of PGMs in the slag. Various feed materials have been treated successfully in furnace campaigns running at power levels up to 1.5 MW. The longest of these campaigns was run continuously for more than four months, treating more than two thousand tons of materials such as low-grade concentrates, revert tailings, and converter slag. The campaigns demonstrated the robustness and versatility of the process, and proved that it is possible to sustainably produce a PGM-containing alloy and discardable slags containing an average of less than 1.4 g/t of PGMs over the entire campaign. Keywords: DC arc, smelting, furnace, platinum, PGM, revert tailings, converter slag, ConRoast

Jan 1, 2004

By E. Kalhan, C. Kincal

One of the biggest problems in open pit mining is slope stability. The avoidance of slope stability problems depends on a comprehensive assessment of environmental and operational factors such as geology, climate, and mining activities. In the light of these considerations, slope geometries should be determined in accordance with engineering principles. In this research, the slope stability of rock slopes in a mine opened in a gneiss geological unit has been studied. Production activities were ceased in 2019 due to the slope movements in the eastern part of the quarry. In order to safely continue the production, field works were primarily carried out. To assess slope stability, detailed discontinuity surveys were carried out on bench faces. Subsequently, kinematic and numerical analyses were employed to identify geometries that are expected to maintain stability under the prevailing geological structures and environmental constraints. Engineering geological studies such as scanline measurements in steps and investigation of potential failure mechanisms by kinematic analysis technique were carried out. Small module reactor and Q-slope rock mass classification systems were used to evaluate slope stability. Based on these classification systems potential unstable rock slopes were identified and thematic maps were created using geographic information systems. The most suitable slope geometries are suggested for these slopes. In addition, the improvement methods suggested by the rock mass classification systems were proposed for these slopes.

Sep 1, 2025

By V. Oberholzer

The presence of fines (particles smaller than 6 mm in diameter) causes hydrodynamic problems in gasifiers and therefore it is of great importance to minimize the amount of fine coal in the feed. This serves as motivation for understanding coal?s breakage mechanisms, which could lead to the ability to predict the generation of fines. The aim of this project was to simulate the pulsating effect of the conveyor belt in order to investigate factors influencing the breakage. Results indicated that an increased initial particle size had an increasing effect in the amount of fines generated. Weathering had an insignificant effect on the breakage of coal. To conclude, a breakage model was developed to describe the rate of breakage out of the top size when a combination of two breakage modes is present.

Jan 1, 2009

Plurigaussian simulation allows constructing lithofacies or rock type models that reproduce the contacts between facies in accordance with the geologist?s interpretation. Its implementation requires inferring the local facies proportions, but the uncertainty in the true proportions is not accounted for. The simpler model with constant facies proportions may not yield realistic results, due to the possibility of obtaining facies at locations where it is geologically unlikely to find them. This article presents a variation of the plurigaussian model, in which the facies proportions are represented by random fields. The realizations can be made conditional to soft geological informati onto account for local changes in the facies proportions. The model is illustrated via a case study of a porphyry copper deposit where four Gaussian random fields are simulated conditionally to drill hole data and to constraints on the probability of finding a given facies at specific locations (control points) in the deposit. Then the first two fields are truncated using the random thresholds defined by the last two, generating a three-facies model. The proposed random proportion model proves to be simple to use and to account for spatial variations of the geological characteristics and for the uncertainty in the facies proportions. Keywords: categorical variable; lithofacies; truncated plurigaussian simulation; regionalized proportions

Jan 1, 2008

By B. Geelhoed

Gy&apos;s formula for the variance of the fundamental sampling error (FSE) is based on a model that assumes independent selections of particles. A generalization of this model, which can theoretically lead to more realistic representation of actual sampling processes, was, for the first time, experimentally verified by repeated sampling from a well-mixed batch consisting of particles of zirconium silicate of two different sizes. The predictions for the sampling variance made in Gy?s theory and in the more general model were compared with the experimentally observed sampling variance. During the experiment performed here, the variance prediction made in the theory of Gy ranges from 0.96 × 10?7(g2/g2) to 0.95 × 10?7(g2/g2), which is approximately a factor 60 lower than the experimentally observed variance (5.76 × 10?6(g2/g2)). The difference between the Gy prediction and the experimentally observed variance cannot be fully explained by the uncertainty in the experimentally obtained variance due to the finite number of degrees of freedom. Therefore, the difference is likely to be a significant effect. This shows that, under the experimental conditions, Gy?s formula for the variance of the FSE probably undervalues the sampling variance, unless an empirical correction factor for grouping and segregation is applied in the theory of Gy. The new theory resulted in a variance prediction of 6.00 × 10?6(g2/g2), which is much closer to the observed variance than the Gy prediction. This indicates that the generalization of Gy?s theory offers an outlook for providing a better variance prediction than the Gy predictor in the case of dependencies between particle selections. The results are discussed. Keywords: sampling, variance, parameter for the dependent selection of particles, Gy, experiment.

Jan 1, 2009

By G. Sahasrabudh, G. DeIulii, K. P. Galvin

A novel agglomeration process has been developed, utilising a concentrated water in oil emulsion as a binder. The binder is typically 95 wt% water, the remainder being an organic phase containing kerosene and the emulsifier— sorbitan mono oleate (SMO) in equal mass portions. The binder can selectively agglomerate hydrophobic particles in less than 7 seconds. The agglomeration has proven to be highly effective in recovering any particle of virtually any size, provided the particle is hydrophobic. The approach has previously been applied to coal, magnetite and silica. In the paper, sodium isobutyl xanthate (SIBX) and its much longer chain version, potassium tetradecyl xanthate (C14X), were used separately as collectors to selectively agglomerate chalcopyrite from the ore. Xanthates attach to metal sulphides, thus hydrophobising sulphide mineral particles. Chalcopyrite is a sulphide mineral; however, it hydrolyses quickly in the presence of water. Ethylenediaminetetraacetic acid (EDTA) was used to remove the oxide layer formed at the surface of the mineral particles after wet grinding, leaving behind a freshly generated metal sulphide surface. After the EDTA treatment, the ore particles were quickly conditioned at an optimised pH as well as collector concentration and agglomerated. The agglomeration achieved a copper recovery of 75% within 7 seconds, consistent with the recovery achieved by flotation over much longer time scales. Keywords: Chalcopyrite, agglomeration, froth flotation, interfacial chemistry, X-ray fluorescence, selective hydrophobisation

Jan 1, 2020

Introduction to high temperature oxidation of metals, by N. Birks and G. H. Meier, London, Edward Arnold, 1983. 19S pp. R30,25. Reviewers: M. B. Cortie and G. G. Garrett Seam gas drainage with particular reference to the working seam, edited by A. J. Hargraves. Woollongong, The Australasian Institute of Mining and Metallurgy. 370 pp. $A30. Reviewer: C. J. Fauconnier New books Solvent extraction Part I, by G. M. Ritcey and A. Ashbrook. Amsterdam, Elsevier, 1983. Elektronishche Steuerungen und Mikroprozessoren im Bergbau. Sicherheit und Zuverlassigkeit Environmental impact control. Philosophical and procedural guidelines for the planning professions Mineral processing and extractive metallurgy. Reagents in the minerals industry.

Jan 1, 1984

By van der Merwem J. N.

"Coal pillars are prone to scaling over time, and are progressively reduced in size and consequently subjected to increased load and reduced strength. The effective factor of safety therefore reduces. Databases of failed and stable pillars cannot, therefore, be regarded as definitive.More than a decade has passed since the first attempt to predict the lifespan of pillars. This paper presents a review based on the new database and pillar strength equations. The methodology for the analysis is based on the fact that pillars scale. The ultimate safety factor (defined as the ratio between strength and load at the time of failure) is derived, followed by an estimate of the rate of pillar scaling, which results in an estimate of the expected time of failure.It is suggested that the predicted lifespan should not be used as an absolute indicator of expected pillar life, but rather as an index – the pillar life index (PLI) – to complement the safety factor and related probability of failure when evaluating pillar stability. A pillar with a predicted lifespan of less than 500 years would be regarded as being in danger of imminent collapse, while a lifespan of at least 1 000 years should be required for long-term purposes.The differences between this analysis and that published in 2003 constitute a strong case for regular reviews of all the empirically based stability indicators, namely the safety factor, probability of failure, and the PLI. Reviews at 10 year intervals appear to be required.The constants in the proposed method are valid only for the Witbank No. 1, 2 and 4 seams and the Highveld No. 2 and 4 seams.IntroductionIt is well known that over time, coal pillars are prone to scaling, and thus are progressively reduced in size and consequently subjected to increased load and reduced strength. The effective factor of safety therefore reduces and it is quite possible for failure to occur after a long time interval. In fact, scrutiny of the database of failed pillars indicates that failure occurred after periods ranging from 2 to 49 years, and not immediately after mining.The factor of safety for pillars at the time of failure is not the same as that immediately after mining. The concept of the safety factor is based on the pillar strength obtained by statistical analysis of pillar dimensions immediately, or within a month after mining, when the surveys are conducted. Strictly speaking, the safety factor as commonly used is valid only immediately after mining, and it continues to reduce thereafter. The ideal situation would be to base analyses of pillar strength on pillar dimensions at the time of failure, but that is simply not possible because the real dimensions cannot be measured at that time."

Jan 1, 2016