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By J Safarian, D Hariswijaya

Treating bauxite residue as an alternative source of metals for iron and aluminium industry is an approach to promote circular economy in metal industries. Reduction of metal oxides with H2-based process is an important step on decarbonisation of metal industry. In this study bauxite residue pellets were prepared and were reduced with different H2-H2O gas compositions at different temperatures which yielded with various degrees of reduction. The bauxite residue pellets were made from a mixture of bauxite residue and Ca(OH)2 powders and sintered at 1150°C. Hydrogen reduction was carried out on the oxide pellets using a resistance furnace at elevated temperatures in controlled reduction atmosphere of H2-H2O gas mixtures which resulted in reduction of iron oxides in the pellets. Unreduced and reduced pellets were subsequently heated to 1400°C to study their smelting behaviour using differential thermal analysis (DTA) and thermogravimetric analysis (TGA) technique to investigate the evolution of phases related to slag formation via smelting. Equilibrium module of Factsage™ ver 8.1 was utilised to analyse results of thermal analysis. It was observed that Tricalcium phosphate-β (CaP2O8), Bredigite (Ca7Si4MgO16), Rankinite (Ca3Si2O7), and calcium alumino ferrite (Ca[Al,Fe]6O10) were formed during thermal analysis as intermediate phases. The initial slag formation for sintered and reduced pellets occurred at 900°C which mainly contains calcium and small amount of strontium component. The slag formation rate increases significantly starting from 1100°C when iron oxides started to form initial molten slag phase which then followed by other oxides dissolution in the system. Gas formation was observed at 1180°C and it was found that the gas to be released from unreduced pellets is O2 meanwhile the gas to be released from reduced pellets is SO2 gas. Gas formation rate for both pellets start to increase at 1280°C when the remaining chemically bonded gas start to be released from the system.

Jun 19, 2024

By G. L. Schroeder, C. H. Summers, D. B. Lindsay

INTRODUCTION The requirement that is placed on owners and operators of underground mines to protect workers against the health hazard of inhaling radioactive materials which are short-lived decay products of 222Rn can be satisfied by applying a considerable variety of what we may call "engineering" solutions as well as a number of "administrative" remedies to the problem. The most obvious of the "engineering" approaches has always been that of forced ventilation, in which relatively clean (i.e., radon-free) air from aboveground is drawn or pushed through the mine workings by a system of fans and ducts. This relatively clean air, in sweeping through the drifts, stopes and haulageways, dilutes the radon and radon-daughter concentrations in the air of the mine, and performs the added beneficial function of removing the daughter-mixture quickly enought to limit grow-in of the longer-lived nuclides in the group that make up the "toxic trio" on which the Working Level (IM) unit is based. Effective as the dilution-ventilation method is for control of WL in most underground mining situations, however, the increasing strictness of control measures that have been imposed on the mining industry over the last two decades have demanded measures of even greater effectiveness. In times of poor markets for yellow-cake and other products of the mines, mine operators are pressed to reduce operating costs, and the installation of additional primary ventilation capacity can be a severe burden on a mine that is already laboring under an unfavorable earning power. When traditional dilution-ventilation systems alone cannot meet the requirement for WI, control, radiation safety engineers and ventilation engineers begin to look at alternatives and auxiliary methods. Since the radon which produces the toxic daughter products originates in the rock of the mine walls, and since, in most United States mines, that rock is a porous sandstone through which air can move under the effect of atmospheric pressure gradients, and through which radon can diffuse relatively freely, one way to help control the growth of WL would be to hinder the escape of radon from that reservior of porous rock. An appealing; method for hindering that natural flux of radon-polluted air from the walls of the mine has long been apparent; namely, to apply a low-permeability coating over the surface of the rock, thus sealing the radon in place and, in theory at least, preventing its escape into the mine air. Our 1970 report to the U.S. Federal Radiation Council on the subject of cost impacts of proposed changes in the occupational standards for exposure of underground uranium miners to airborne radon daughters noted the possibility of using polymeric wall sealants as a means of controlling radon-pollution of mine air. Since that time a number of reports have appeared in the technical literature advocating this technique for restraining the escape of radon from building materials, mill tailings, and other materials containing 226Ra, in addition to the surfaces of underground mine workings. During this period, some controversy has occurred over the question of the probable effectiveness of wall sealants in limiting the escape of radon from the rock. Our 1970 report speculated that flaws (cracks and "pinholes") in the coating might be all but unavoidable in practice, and that even a conservative estimate of the frequency of such flaws would lead to a prediction of ineffectiveness. Hammon et al, in a laboratory evaluation of radon sealants conducted by Lawrence Livermore Laboratory of the University of California in 1975 on behalf of U.S. Bureau of Mines, concluded that a wide variety of polymeric coatings would provide "nearly 100% effectiveness" in restrain¬ing escape of radon from mine wall surfaces if applied in "thicknesses between 5 and 10 mil" (125-250 [y]pm). John Franklin and co-workers at the U.S. Bureau of Mines laboratories in Spokane, Washington, have carried the experiments with polymeric sealants through additional laboratory tests and into actual mine environments, reporting that selected sealants could provide attenuation of radon flux by a factor of four (75-80% reduction). Robert Bates and John Edwards of USBM developed a computer-assisted mathematical/physical model that predicts a relatively small effect of flaws in a low-permeability coating on the radon flux from a sandstone-type matrix. FIELD TEST Since all actual experimental work with wall sealants showed some beneficial effect on radon attenuation (even if not as exciting as the "nearly 100%" predicted by Hammon), USBM was encouraged to extend its evaluation to an actual operating uranium mine, and awarded a contract for that work to Arthur D. Little, Inc. in September 1979. We were fortunate in obtaining the voluntary cooperation of Atlas Minerals Division of Atlas Corp., who operate a mill and several underground mines in and around Moab, Utah. Atlas made available for our use a small T-shaped drift in their Pandora Mine in LaSal, Utah, and provided space for instrumentation and recordkeeping by our field crew in a surface building near the mine entry. Atlas also provided electricity and water to the test site, together with assistance in establishing necessary ventilation, removing rubble from the site, conducting periodic WL surveys and furnishing auxiliary man-power for the heavy hard work of coating the walls with gunite prior to application of the polymeric sealant. The generous coopera-

Jan 1, 1981

By D R. Seaman, Y Peng, F Burns, D J. Bradshaw

Telfer Gold Mine is a gold-copper mine in the eastern Pilbara of Western Australia, owned and operated by Newcrest Mining Limited. TelferÆs ores are processed by a dual train, 21 Mt/a, concentrator where gold is primarily recovered by gravity concentration and the flotation of a gold-bearing copper concentrate. A minor portion of TelferÆs gold is also recovered to a pyrite concentrate that undergoes treatment via carbon-in-leach (CIL) cyanidation. The gold recovery of the CIL circuit is below the industry standard and operates with a high reagent consumption rate.With a new fine-grinding facility, the Telfer process plant will have the ability to liberate copper and gold from the pyrite concentrate and recover these by flotation prior to the existing leaching circuit. This process will improve overall gold and copper extraction, while reducing reagent consumption in the CIL circuit.This article is an overview of the development of this methodology. Size-by-size analysis and mineralogical studies of monthly composites identified factors limiting the CIL gold recovery and the causes of high reagent consumption. Laboratory studies confirmed a relationship between the regrind product size and gold leach recovery. The addition of an intermediate flotation stage ensured that the regrind circuit will improve overall recovery while at the same time reducing operating costs. Without the addition of a flotation stage, the additional reagent consumption from regrinding alone would be prohibitive towards increasing recovery. Pilot scale studies confirmed the conditions required to selectively recover gold and copper from the pyrite concentrate as well as establishing the full scale regrinding energy requirements. Construction of this facility at Telfer is currently underway.CITATION:Burns, F, Seaman, D R, Peng, Y and Bradshaw, D J, 2012. Development of a pyrite regrind circuit at Telfer Gold Mine, in Proceedings 11th AusIMM Mill OperatorsÆ Conference, pp 45-52 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Oct 29, 2012

By J. B. Gwiazda

Maintenance of a constant load setting throughout longwall support units and selection of the proper initial bearing capacity depending on the type of roof strata are the basic factors that ensure good performance of roof support in a longwall. These requirements can only be met by hydraulic support. The greatest advantage of hydraulic support is achieved when uniform pressure is imposed on the roof throughout the length of the longwall. Such support, however, is provided only if each of the support units acts with the same force against the roof, i.e., has the same setting load. In such cases, the roof behaves like a uniform plate without bending and shearing stresses, thereby ensuring an undisturbed structure. Without a positive set value system, achieving an equal setting load for all units of the longwall support is impossible. Due to alterations of the feed line pressure of support units as well as some reasons related to man's psychology, operators extend the height at different setting loads. This produces nonuniform roof stress, and disturbs the structure. Consequently, the roof usually cracks. The author has developed a positive set value system, which is described in this Technical Note. Selection of the setting load Two pressure values in the feed lines are usually applied in longwall hydraulic systems. Lightweight support is fed by 25 MPa (250 bar) liquid, while heavy duty units receive a nominal pressure of 31.5 MPa (315 bar). Such pressure is required not only for the props but also to power the adjust¬ment jacks and the advancing ram. If the feed pressure is too low, there will be difficulty in shifting the unit despite the inversion system of the advancing rams. On the other hand, for many roof types, the feed pressure often appears to be too high when applied as the setting load pressure. An excessive setting load acts too strongly against the roof, crushing weak strata close to the roof. The author has recognized a case where an excessive setting load destroyed not only the nearby roof strata but also the strata above a 2-m (6.6-ft) sandstone layer. In addition, an excessive setting load relieves the side¬walls, increasing the resistance when using cutting machines. As a result, the yield of coarse coal is diminished, and increased fines dominate in the final product, lowering its economic value. As indicated, selection of the proper setting load, depend¬ing on the mining and geological conditions of the extracted seam, is extremely important. In some mines, measures applied to prevent disturbances include reduction of the feed line pressure by adjusting the feed pump valve. The disadvantage accompanying reduced feed line pressure is more difficult operation in advancing the ram. Due to the reduced feed line pressure, the force of the advancing ram is much lower than the designed value. Other designs suggest using a third feed line. However, installation of supplementary valves on the support units is required, a time-consuming and expensive procedure. The disadvantages of the powered supports are eliminated by a system designed by the author. So far, such a method of setting load control has not been used in any type of support. Setting load control unit The designed positive valve set for prop loading and the setting load control correspond to existing control systems for hydraulic powered support. The layout of the unit connected to the hydraulic prop control is presented in Fig. 1. The unit is marked LIDS. It incorporates three valves that may operate separately or connected. Valve A automatically opens and closes with liquid flow in the prop feed circuit. The valve is opened when the canopy touches the roof and closed when the support unit is withdrawn. Valve B serves as the setting load control. Valve C automatically opens and closes the flow in the line connecting the under-piston space of E to the prop F with the separator G. The valve block of each support prop is marked BZ. The UDS unit is connected by the hydraulic lines to the F prop control circuit. A valve is connected by H to pressure line J and by K to the G separator. In the UDS-3 version, line L is connected to M, linking the over-piston space of F prop with the G separator. Valve B is linked with valve A by a connector; it is also connected to the under-piston space E of prop F by line P. Valve C is fixed between lines K and P, connecting space E of prop F with the separator G. When setting the support, the liquid flows from line J through separator G, the BZ valve, and valve C to the space E of prop F. When reaching the roof with the canopy, valve A is opened and C closes. In this way it is impossible for the operator to cause the liquid pressure in space E of prop F to reach the level of line J. The prop pressure is set by valve B of the UDS unit. When withdrawing the support, valve C is automatically opened and A closed. Three UDS units have been fabricated and are designated

Jan 1, 1990

By R. Howard

"In projects involving resource development and plant optimization, the use of in-mine geophysics can deliver considerable benefit. The value proposition lies in using very high resolution downhole measurements to model process-plant outcomes. The calibration procedure used to convert in-situ physical properties into process engineering parameters is illustrated using two iron ore projects. Data correction was related either to orebody structure or mineral formation. The deliverable from this approach is a series of high-definition sections indicating the variability of ore delivered to a front-end separation plant and parameters for mine scheduling and stockpile design needed for maximum product yield.RÉSUMÉ Dans des projets de développement des ressources et d’optimisation de l’usine, l’utilisation de la géophysique minière peut engendrer des bénéfices considérables. La valeur de la proposition réside dans l’utilisation de diagraphies à très haute résolution afin de modéliser les résultats de l’usine de traitement. La procédure d’étalonnage utilisée pour convertir les propriétés physiques in situ en des paramètres d’ingénierie de procédés est illustrée en se basant sur deux projets de minerai de fer. La correction des données porte sur la structure de la masse minérale ou de la formation minérale. Les livrable attendus de cette approche sont une série de sections à haute définition, signalant la variabilité du minerai livré à une usine de séparation de première ligne, des paramètres pour l’échéancier de la mine et la conception des haldes requis pour une production optimale."

Jan 1, 2016

By Luis Llaque, Alain Revilla

Full Potential (FP) es un programa global de Newmont con el objetivo principal de proporcionar una mejora sostenible al flujo de caja sin comprometer el valor presente neto de la vida de la mina (Life of Mine “LOM”). Como sugiere el nombre del programa, el objetivo es que la mina donde se implementa el programa opere junto con su sistema de activo a su "máximo potencial". Como programa global, FP es coordinado por el equipo de FP Corporativo, utilizando las diversas herramientas y metodologías de implementación asociadas con el programa.

Sep 29, 2023

Statistical summary of the mineral industry world production, exports and imports 1964-1969 Published by Her Majesty's Stationery Office, London, on behalf of the Institute of Geological Sciences (1971) The Statistical Summary contains a wealth of care fully garnered information on the production, exports and imports of the principal minerals and mineral products throughout the world for the period 1964-1969, and it should therefore be of particular interest to economic geologists and others concerned with marketing and selling minerals. In view of the intractable nature of the source data it is quite an achievement to have published figures for 1969 so expeditiously. On picking up the Summary this reviewer instinctively compared it with the Annual Yearbook published by the United States Department of the Interior. The Yearbooks have been issued annually since 1882, and today each consists of a set of three separate books containing five volumes. Apart from abundant statistics on the domestic and international mineral industries, the Yearbooks are known for their perspicacious reviews and articles on a host of different facets of the world's minerals industry. Physically the Summary is a much slimmer publication than the Yearbook, and it contains no commentary whatsoever other than occasional footnotes. The principal contents are a series of tables showing production, exports and imports of fifty-nine of the world's more important minerals and mineral products. Each table shows the weight of the material produced or traded, by countries, with figures for each of the six years in adjoining columns. In some cases (e.g. diamonds) values in British currency are quoted in addition to weights. Most of the data have been abstracted from national reports and returns, although in some instances estimates are quoted where national sources are silent--as for example the production of platinum in South Africa. One wonders in passing if any useful purpose is served by the continued suppression of this particular figure. The four South African producers must surely be able to estimate each other's production fairly accurately by now, and one imagines that the national export figures can likewise be synthesised. This reviewer has two minor criticisms of an otherwise praiseworthy publication. Firstly it seems an unnecessary anachronism to perpetuate the split between 'Commonwealth' and 'Other' countries-the split can have little practical significance today. Secondly, what is one to make of the statement that the 1966-1969 figures for exports from Rhodesia of chrome ore and copper, to take two examples only, are 'not available'? It seems improbable that no estimates of these figures have been made since the imposition of sanctions, and inevitably one suspects that the figures have been excluded on policy grounds. Hopefully this problem will have been solved when the next Summary is published. M.J.M. South Africa, land of Challenge by Maurice Tyack France Interpresse, 323 pages, price R18-00. Is South Africa God's richest acre? In its mineral wealth it certainly is, but what of the many other complex overlays of emotions, traditions and racial problems? South Africa, Land of Challenge, provides the background. This is reported to be the first comprehensive documentary work of its kind on South Africa. It is a handsome volume measuring 12ft by 9f', is illustrated by at least 600 photographs of South Africa, more than half of which are in colour, and contains 21 original maps and charts. The book covers the full spectrum of South Africa's very existence, its geography, pre-history, its customs, countryside, natural resources and its cities. The more controversial categories are thoroughly discussed, the people, their policies and politics. Mr Tyack is eminently qualified to write on South Africa, being the author of over 20 books on other African countries and having spent more than three years of research, in the field, in South African archives and also in various libraries and museums, in preparation for this book. The book claims to be free of any party political bias and appears to be an honest attempt at a factual survey. However, a work of this magnitude will undoubtedly leave an impression on the reader and the tone of this book is decidedly patriotic. This is a useful reference work on South Africa that most South Africans will be proud to own. J.P.H. Ergonomics and physical environmental factors This publication by I.L.O. of a symposium in Rome in September 1968 has much useful information for the mining industry in regard to the health and productivity of its labour force. The first subject dealt with is mechanical vibrations. From this section it is clear that men who drive, for many hours each day, vehicles which vibrate in the 6 Hz range are subject to serious physiological and psychological effects. The former may show as changes to the vertebral column (which has a resonance in the 4-5 Hz range). The circulatory system is adversely effected by values in the 6-10 Hz range and the vision is impaired at between 40-100 Hz. Less definite knowledge exists of the effects of vibration on the hand-arm system from mechanical drills etc, although Raynauds disease of the blood vessels of the hand and degenerative diseases of the elbow and shoulder joints occur in men using drills which vibrate, respectively, in the higher and low frequency ranges. Industrial noise is extremely well dealt with and the latest knowledge is summarized. Clear information is given on the relationship between time of exposure to

By P. J. A. Beukes, C. C. La Grange

INTRODUCTION Some years ago the authors proposed a modified procedure1 for carrying out micum index determinations2. The modifications consisted of using a drum of 50 cm internal length, i.e. half the length of the standard micum drum (a modification also adopted by the B.S.J.3), and the use of a + 25 mm coke instead of + 60 mm coke for the test. The proposed modified test and formulae for converting results obtained from it to standard micum test results and vice versa were based on the results obtained during the investigation of about 200 test cokes made in ovens of the South African Steel Industrial Corporation Limited (Iscor), the Fuel Research Institute co-operating with the experiments. Extensive testing of cokes has since been done at the Fuel Research Institute using both the standard and the modified procedures. The numerous test results obtained during this trial period fully support and strengthen the confidence originally expressed by the authors in the modified testing procedure. The scope of the modified micum index determination has in the meantime been extended by applying the testing procedure also to test cokes made in the Institute's experimental coke ovens, and also in this case the modified procedure has proved its usefulness. A difficulty is that when comparing the characteristics of cokes made in the Institute's two experimental coke ovens with each other and with those of cokes made from similar blends in commercial coke ovens, for example those of Iscor, the index values obtained invariably do not check exactly due mainly to inherent differences between the different types of oven4. For instance, resistance to abrasion of a coke made in Iscor's ovens is always appreciably higher than that of a coke made from the same blend (and having a similar moisture content) in the experimental ovens. This makes direct comparisons and the prediction of quality to be expected from commercial products, based on experimental coke oven results, difficult. It may also be added that workers in Great Britain some years ago reported a similar experience5. In the authors' opinion the main reason for the discrepancy is the relatively low height of the experimental ovens resulting in a much lower static pressure on the charge during coking than that obtaining in a commercial oven. The difficulty was pointed out and briefly discussed in a publication4 which appeared at a time when the micum test for coke evaluation had only just been introduced at the Institute, so that only B.S. shatter and abrasion test results were available for discussion in the publication. Needless to say, the introduction of the micum test by no means solved nor even alleviated the problem. On numerous occasions during the past few years the Institute has had to conduct contract investigations in its experimental coke ovens on behalf of companies which invariably desired an indication of the quality of coke likely to be obtained from certain coals or blends when coked commercially, more particularly in Iscor's type of coke ovens. It was, therefore, essential to accumulate statistical data which would enabe the prediction to be readily made. Fortunately the compilation of comparative coke-quality index values for the three types of oven under consideration has become possible as a result of coking investigations conducted in co-operation with Iscor over a number of years. During this period Iscor, on numerous occasions, kindly made available to the Institute portions of coals and blends coked in the Corporation's coke ovens, thus enabling parallel coking tests to be carried out in the Institute's experimental ovens. In this way enough statistical data became available to establish relationships between coking results obtained in Iscor's ovens and results obtained on similar coal charges in the Institute's ovens. The most practical and acceptable method of overcoming the problems of finding a suitable and simple basis enabling direct comparison between the cokes of the different origins mentioned to be made, is the assignment of descriptive arbitrary quality ratings, as indicated below, to the cokes obtained, it being argued that the quality rating assigned to coke made from a given coal or blend in the different coke ovens under standardized conditions should be the same, irrespective of differences in the index values obtained when subjecting the cokes to the usual coke evaluation tests. The system, which should enable the sponsors of coking investigations in the Institute's experimental ovens to make their own interpretations of the results obtained, is explained below. ARBITRARY QUALITY CLASSIFICATION OF COKES The Ml0m index1 (percentage material smaller than 10 mm after the test) of a blast furnace coke is a measure of its tendency towards breeze formation during handling -a low value indicating a low tendency. This is generally of greater importance under ruling South African conditions, than the M'40 index1 (percentage material larger than 40 mm after the test), which can be regarded as a measure of resistance to shatter. In fact, if the M10m value of a coke as made in Iscor's ovens is higher than about 11, the tendency of the coke to form breeze during handling and in the blast furnace is relatively so high that hardly any notice need be taken of its M'40 value, even if this is comparatively high. For this reason, therefore, much more weight should normally be assigned to Ml0m indices in the quality evaluation of cokes made from South African coking coals and blends, as the coals

"Recognising that the mining industry is wedded to the break-even cut-off grade rather than a cut-off grade that optimises value, the author examines some perils and pitfalls in the use of break-even grades that lead to what the author calls ‘negatively geared ore reserves’.Negatively geared ore reserves occur when a portion of the ore reserves actually has a negative value and is effectively being subsidised by the more profitable portion of the ore reserves. This situation is generally difficult to detect with the level of information available to the public investor. The ore reserves can still be quite profitable in total, but now possess a level of risk that may not be appreciated.The risk profile (financial risk) of a negatively geared ore reserve is far greater than for a non-negatively geared ore reserve. Like a negatively geared investment property in a falling property market (for readers familiar with the Australian tax minimisation investment strategy), when the price of a commodity falls, a company that possesses a negatively geared ore reserve loses value at a magnified rate relative to a company that has a neutral or positively geared ore reserve. In the extreme case of a major commodity price reduction, as happened with the gold price in mid-2013, the negatively geared ore reserve can easily become negative value in total, not just in part.The negatively geared ore reserve is sometimes intentional – for instance in order to increase published ore reserve tonnages and meet the CEO’s key performance indicators – but is also often unintentional, resulting from various errors in the break-even calculation. Common errors in calculating a break-even grade include using fixed metallurgical recoveries, not including allowances for ideal laboratory versus real plant results, ignoring available operational data and using modal rather than mean parameter values in cost and performance calculations.This paper examines a common value distribution profile in many ore reserves and the effect of common break-even grade errors on this profile.CITATION:Poniewierski, J, 2016. Negatively geared ore reserves – a major peril of the break-even cut-off grade, in Proceedings Project Evaluation 2016, pp 236–247 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Mar 8, 2016

By E Donskoi

In order to develop downstream processing routines for iron ore and to understand the behaviour of the ore during such processing, extensive mineralogical characterisation is required. To determine mineral associations, mineral liberation and grain size distribution, imaging techniques have to be used. There are two main imaging techniques used for the characterisation of iron ore, ie, Optical Image Analysis (OIA) and Scanning Electron Microscopy (SEM). In this article a QEMSCANTM system is used as an example of SEM methodology and results obtained from it are compared against results obtained by the CSIRO ‘Recognition3/Mineral3’ OIA system.Both OIA and QEMSCANTM systems have advantages and drawbacks. Even though the latest QEMSCANTM systems can distinguish between major iron oxides and oxyhydroxides, it is still problematic for QEMSCANTM to distinguish between iron ore minerals very close in oxygen content, eg haematite and hydroheamatite, or between different types of vitreous goethite.QEMSCANTM also can misidentify minerals with close chemical composition, ie haematite as magnetite and vitreous goethite as haematite. In OIA, iron minerals with slight differences in their oxidation or hydration state are more easily and directly recognisable by correlation with their reflectivity. In both methods, the presence of microporosity can result in some misidentification, but in SEM methods misidentifi cations due to microporosity can be critical. Low resolution during QEMSCANTM analysis can significantly affect the textural classifi cation of particle sections.The main conclusion of this study is that, for low iron content ores or tailings, QEMSCANTM can provide much more detailed information on the gangue minerals than OIA. However, for routine characterisation of iron ores with high iron content and containing a variety of iron oxides and oxyhydroxides, OIA is a faster, more cost-effective and more reliable method of iron ore characterisation. A combined approach using both techniques will provide the most detailed understanding of iron ore samples being characterised.

Jul 11, 2011

By Juan J. Monsalve, AMAN SONI, Nino Ripepi, Richard Bishop

The design and strength of pillars are important parameters while implementing the room-and-pillar mining method in a limestone mine. The process becomes complex with the presence of the karst voids which are frequent phenomena in a carbonate rock formation. The karst system is a network of cavities that along with an omnipresent fracture network may lead to ground control issues in a limestone mine. The objective of this study is to provide a safe and reliable alternative design for a pillar in a case study mine for safe and efficient ore recovery. This research is a continuation of a previous study aimed at studying the effect of karst void on the strength of the same limestone pillar. It was assessed from the calculated pillar strength and the estimated redistributed stresses that the pillar is capable of countering more stress than is currently imparted on it. Numerical modeling studies combining in situ stress determination and excavation sequencing were used to estimate the existing stresses on the pillar. Different possibilities of modification to existing pillar design were simulated using Distinct-Element Modeling (DEM) to determine a better design. Discrete fracture networks (DFNs) were included to imitate the effect of the presence of discontinuities. An effort is also made to understand the variation of pillar strength with height for karst-affected pillars using two different constitutive models. This study serves as a preliminary investigation in terms of improving the future design of pillars and also while recovering old pillars prone to karst formation.

Jan 30, 2023

By K Edwards, A J. J Mouton

Hematite represents the vast majority of iron ore exported from Australia as the development and processing cost of magnetite is generally seen as a barrier to the development of vast magnetite resources. However, a number of magnetite projects have attracted a significant amount of interest and investment from large Chinese steelmakers.This paper compares the relative merits of the development of Australian magnetite and hematite ore from the point of view of the steelmaker, taking into account the factors that influence the capital costs of the development of the ore types and the relative costs of operating the process plants to treat these ores. By modelling the relative operating cost for the production of hot rolled coil in China for the two ore categories, the paper demonstrates the economic drivers behind the Chinese steelmakers’ large investments in the magnetite resources of Western Australia. It does this by estimating the potential value of magnetite concentrate to the steelmaker in the event that the maximum potential is realised by operating the blast furnace with a 100 per cent pellet burden.Under the conditions described, it was found that direct reduction (DR)-grade magnetite concentrate could command a price premium over 62 per cent Fe fines of up to 46 per cent (December 2014 on a cost and freight (CFR) basis). The corresponding premium for blast furnace (BF)-grade magnetite concentrate is 17 per cent, which demonstrates the dependency on grade.The capital expenses for a 14 Mt/a concentrate plant is estimated at US$3.1 B for BF-grade magnetite and US$3.35 B for DR-grade, with operating expenses at US$45 and US$50 respectively. The corresponding free on board price for a break-even net present value is calculated at US$82/t for BF-grade magnetite and US$90/t for DR-grade. The minimum long-term iron ore price (62 per cent Fe) that would be required to underpin such a project is estimated at US$78/t CFR for BF-grade magnetite and US$68/t CFR for DR-grade concentrate.Apart from the financial benefits, it is demonstrated that the use of magnetite concentrate in ironmaking has the potential to reduce CO2 emissions by approximately 400 kg per tonne of hot rolled coil.CITATION:Mouton, A J J and Edwards, K, 2015. Magnetite – a higher-grade blast furnace feed and its potential benefits for the ironmaker, in Proceedings Iron Ore 2015, pp 613–620 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 13, 2015

By A J. Ryan

The Ernest Henry copper-gold Mineral Resource is the major asset heldby Ernest Henry Mining Pty Ltd, comprising A$6.4 billion of containedmetal. The case history presented below describing the evolution andevaluation of this resource generally observes best industry practice and is a good example of effective due diligence in the resource delineation and estimation process. Early reconciliation data from mine production confirms the accuracy and validity of the resource estimate, providing valuable assurance to the mine owners.

Jan 1, 1998

By C. Langer

Although the existence of precious metals, particularly those of the platinum group, in the nickel ores in the Sudbury district has been known for a good many years- almost as long as the ore deposits themselves- the Mond Nickel Company were the first to effect their recovery on a commercial scale. The precious metals values of the different Sudbury ores vary considerably, and are generally higher in the copper-rich ores. These values are so very small that a direct determination in the ores gives quite unreliable results. This is not surprising if one considers that some of the ores contain on1y one unit of total precious metals in about two million units of ore, and that this one unit will consist of seven different metals, each with a different chemical behaviour. It is only after the ores have been smelted to a Bessemer matte containing about 80 per cent combined copper and nickel that the precious metals content can be determined with some degree of certainty. From the figures thus obtained, the precious metals content of the ores can be calculated, provided the tonnage of ore used to produce one ton of matte is known. The recovery of these minute quantities of precious metals out of ?such a large amount of gangue and sulphides of iron, copper, and nickel seems at first glance a commercial and technical impossibility. Nevertheless, science has found a way to conquer all the difficulties, and quite substantial quantities of the precious metals contained in the Sudbury nickel-copper ores have been regu1arly recovered for many years by the Mond Nickel Company.

Jan 1, 1927

The history of the rail and mining industry in Queensland over the past century is inseparable. Indeed it was mining that led to the construction of the line west from Townsville and that shaped its current alignment. The discovery of gold at Charters Towers started construction of the Great Northern Railway west from Townsville in 1879, with completion to Charters Towers in 1882. The line was extended to Hughenden in 1887 and to Cloncurry in 1907 - driven by copper mines both to the south west and the north west of Cloncurry. The mining industryÆs association with funding the construction of rail lines to service their individual mines began in this area when the Hampden and Mount Elliot mine companies agreed to pay 50 per cent of the cost of the line from Cloncurry. The line to Selwyn was completed in 1910. Copper mining saw the construction of many other rail lines in the Cloncurry area in the 1910 - 20 period before copper prices slumped. Included in these lines were the Malbon to Duchess line and the line to Kajabbi and beyond.

Jan 1, 1998

By Swades Kumar Chaulya, Vikash Kumar, Sushma Kumari, Naresh Kumar, Girendra Mohan Prasad, Monika Choudhary

Opencast mines are vulnerable to inclement weather conditions as it disrupts the day-to-day activities at these mines. To address the problem, a perceptive driving assistant system (PDAS) has been developed to increase the visibility of real-time video of the road in front of the vehicle for operators of heavy earth-moving machinery (HEMM) at opencast mines in harsh weather conditions.

Apr 1, 2023

By J. Ouellet

The design of hydrostatic liner for a deep potash mine shaft in Saskatchewan poses tremendous technological challenges. Technology has evolved to a point where one could consider a new innovative approach to hydrostatic liner design. But there is still a place for conventional technologies such as cast iron tubbings. As will be discussed in this paper, improvement to manufacturing process, installation procedures and even new materials can bring this technology to an excellent level of performance. With the spur of new projects in the potash industry of Saskatchewan, the authors worked in various shaft liner designs from the conceptual stage through to the construction phase. They were either directly leading the design work on some of these projects or acting as an external reviewer on others. Most of these shaft liner designs involved cast iron tubbing technology and ground freezing. The major challenge in sinking potash shafts in Saskatchewan is to go through a water-bearing ground formation under high pressure, known as the Blairmore or Manville fm. The only proven stabilizing technology for these formations is ground freezing. For a potash shaft, dry shaft conditions are highly desired, so a fully waterproof hydrostatic liner is required. All the existing potash mine shafts in Saskatchewan, except for two with composite steelconcrete lining, used the cast iron tubbing technology. Two new shaft tubbing liners have been constructed recently, one in the Athabasca Basin and one for a potash mine in Saskatchewan. When the authors began working on these new projects, what became quickly evident is that the technological knowhow for designing and constructing tubbing liners was no longer available within Canada and that the past construction methods used to sink tubbing lined shafts was obsolete. The tubbing liner for the shafts sunk in the sixties were designed either by German or British companies. In both of these countries only few tubbing lined shafts were sunk in last 40 years and the technology used in the sixties was not improved. There are only few countries where tubbing is routinely produced and installed: China, Russia, Ukraine and Poland. Out of these countries Poland is the one where this technology is the most advanced. The authors, working in close collaboration with engineers from Poland, attempted to rebuild this knowledge for the mining industry in Canada. This process involved thorough analysis of old technical reports, literature review and study of current polish design and construction practices as well as site visits of existing shafts. Through actual construction experience, some lessons were learned as well as ideas for improvements. In this paper, the authors will present a succinct review of design and procurement process. But, the practical aspects of the construction and installation methodology will be treated with more emphasis, more specifically the practice leading to good liner performance. Some of the poorly understood technical aspects of tubbing installation will be discussed more thoroughly and some currently accepted myths dispelled.

Aug 1, 2013

By D Kilkenny, D Nardini, P Kline

This paper is a case study highlighting learnings and success in two areas challenging underground miners: change management and collaboration. It examines the implementation process of a new automated loading system in the bulk stoping area of Telfer Mine in Western Australia, controlling multiple load haul dump machines. Here, the new technology has brought together different functions and different companies to deliver a well performing automation system in a changing and challenging mining environment. The introduction of a new and different technology on site has driven collaboration between all stakeholders to break down the traditional barriers to achieve the sustained performance required. Historically, different teams have operated within silos and not always aligned to deliver a ‘best for project’ solution. Technical teams focus on the mine design, maintenance teams focus on reliability and operations execute the plan. Mining contractors focus on productivity and cost. Manufacturers focus on supply of goods and service. Optimising each silo individually is not guaranteed to provide an optimal mining solution. Involving elements from each team as seen at Telfer, has resulted in an optimal mining solution for the automated loading system. A change of focus from the technical teams to include automation parameters in the mine design; operability and maintainability built into the system design allowing for collective ownership and responsibility resulting in a more resilient solution. The implementation of the new technology was not only a technical change but more importantly, successful management of change. Elements of successful change management such as communication, vision, leadership and planning, are essential to harness a trusting site culture which is responsive to new technology. Learnings and challenges from this change process will also be discussed with the intent to improve the success rate and timelines of technology implementations in the future.

Mar 15, 2021

By N. Aslin, R. Clark, G. Heferen

The methods of harvesting copper from stainless steel mother plates has been investigated from fundamental and operational experience perspectives. At least five different harvesting techniques have been identified that have different effects on product quality, current efficiency, current distribution, growth morphology and ease of stripping. Each method is examined from fundamental principles to understand what occurs during the harvesting process and why the particular effects are manifested. Experience from plants operating the harvesting method is provided to show that the fundamental understanding is relevant. Best practice harvesting method and harvesting management are described.

Jan 1, 2014

By J Franz, S Bock, N Hentrich

Artificial ground freezing is the unrivalled method for groundwater control during mine shaft sinking. Moreover, ground freezing may also provide temporary ground support. The fast development of fibre-optic temperature monitoring devices, three-dimensional freeze pipe survey methods and numerical modelling software provide grounds for developing new and reliable tools for ground control during freeze shaft sinking. This paper presents examples of predictive numerical models that are based on back-analysis and calibration of temperature data that covers the complete construction cycle from pre-freezing through to artificial thawing. The back-analyses showed that standard numerical models are not capable of predicting the prevailing thermo-mechanical conditions satisfactorily. This paper describes the development and implementation of new tools for improved consideration of concrete hydration and latent heat, along with the introduction of other algorithms that improve and simplify the efforts of managing complex numerical models. A function is presented that allows for an automated correction of the model’s mesh based on results of three-dimensional freeze pipe surveys. This paper demonstrates that the latter consideration is important, because the often employed assumption of idealised freeze circle geometry in numerical models impacts on the accuracy of such model’s predictions. Subsequently, the verification of these extended numerical models is demonstrated. Due to its flexibility, the FLAC3DTM code was employed for the numerical modelling. This software provides the possibility of implementing additional functions written in internal FISH or Python language into the models. It is concluded that state-of-the-art techniques for freeze hole drilling and surveying, on-site monitoring and advanced numerical analysis provide a significant potential for both the design and the operation of largescale ground freezing measures. The application of the extended numerical models for predictive simulations during construction allows for temporary shaft lining optimisation, for assisting in operational decision making and for compensating temporary loss of monitoring data. Suggestions are made for required future work in order to further improve the developed ground control tools. CITATION:Bock, S, Franz, J and Hentrich, N, 2018. New tools for ground control in freeze shaft sinking, in Proceedings The Fourth Australasian Ground Control in Mining Conference (AusRock), pp 393–406 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 30, 2018