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By G Nicholas, S Coward, J Vann, R Wolff, S Jackson, S Moayer, A Bye

"A new approach to mineral project evaluation is described called Scenario-Based Project Evaluation (SBPE). SBPE incorporates well established æscenario planningÆ and æscenario thinkingÆ approaches, which are used extensively in other industries,?most notably the?petroleum industry. While most scenario planning is qualitative and non-probabilistic, the approach described here includes both quantitative and qualitative aspects. SBPE is complementary to optimisation approaches and allows mining firms to build and operate projects that are more robust in the face of adverse changes as well as proactively respond and exploit the upside offered by a realistic assessment of variability. The approach has potential to improve project performance and decrease operational risk. It also has important applications for currently operating mines.Conventional evaluation of metalliferous mineral deposits uses estimates of metal grades and deleterious attributes in singular æorebody modelsÆ, usually 3D block models, which fail to capture two critical aspects that underpin the realised value of mining projects: the inherent uncertainty of the estimated attributes the spatial variability of attributes, at the scale of mining selection, which impact on downstream processing steps and fundamentally drive realised value. While singular models are appropriate for estimating block averages, conditional simulation (CS) models of the relevant attributes are needed to represent uncertainty and capture spatial variability. Any attributes that can be either built into CS models and/or modelled in the envisaged mining/processing steps, can be evaluated fully through to highly ægranularÆ (ie detailed and un-smoothed) cash flows. A complete view of SBPE relies on selected strategies being tested against plausible æscenariosÆ of factors external to the project and thus outside management control. æExternalitiesÆ captured in such scenarios include market prices, costs and taxes. Via SBPE, project managers may demonstrate that there is material value in preserving or purchasing options that provide future flexibility to respond to the changing externalities or changes in orebody performance. Such flexibility generally comes at a cost but is very difficult to value. It is argued that full value chain evaluation using SBPE can help to value these options for a mining company. A case study for a gold mine is presented.CITATION:Vann, J, Jackson, S, Bye, A, Coward, S, Moayer, S, Nicholas, G and Wolff, R, 2012. Scenario thinking - A powerful tool for strategic planning and evaluation of mining projects and operations, in Proceedings Project Evaluation 2012, pp 5-14 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

May 24, 2012

By T Roberts, D Faulkner, G Watt

Traditional mechanised resin rock bolting in hard rock underground mines requires the installation of resin capsules into a pre-drilled hole prior to the rock bolt being spun into a hole to mix the resin and set the bolt in place. In poor rockmass conditions, especially at depth or in highly altered faults and shears, the condition of the holes often deteriorates as the poor ground around it unravels and falls into the hole. This results in partially or fully blocked holes that either damages the resin capsule when they are being inserted into the holes or prevent it entirely until the hole is continuously cleaned of debris. This deterioration of the hole condition, and continuous need to clean and re-drill holes, also increases the diameter of the hole which in turn affects the continuously mechanically coupled anchorage mechanism of the rock bolt to the rockmass as there is simply not enough resin in the hole to fill the annulus between the bolt and the rock. This results in a large increase to critical heading cycle time and, more importantly, results in the rock reinforcement not performing as per design specification and increases the risk of instability. To overcome the quality control issues with resin bolt installation in poor ground, a single pass drill, install, inject and tension self-drilling resin bolt application was designed and implemented for use in such conditions. This system allows the operator to inject a two part polyester resin through a self-drilled hollow bolt to mix and fully encapsulate the rock reinforcement. The self-drilling bolt allows for post injection of the hole after the bolt has been installed into the rock and gives the operator manual or automated control on the amount of resin used in the hole to ensure complete encapsulation and continuously mechanically coupled anchorage to the rock. This allows the reinforcement to work as designed and provide the Geotechnical Engineer extra quality control confidence in these conditions. The implementation of this single pass resin bolting system resulted in significant decreases to the resin bolting cycle time and rock bolt quality control failures. In this paper the design, implementation and embedment into the development cycle will be discussed and the in situ quality control checks used to validate the increase in rock bolt installation quality will be shared. CITATION:Watt, G, Roberts, T and Faulkner, D, 2018. Single pass drill, install and inject self-drilling resin bolt application in poor ground, in Proceedings The Fourth Australasian Ground Control in Mining Conference (AusRock), pp 323–343 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 30, 2018

Driefontein Gold Mine together with Kloof Gold Mine and Beatrix Gold Mine forms part of Gold Fields South Africa. Driefontein Gold Mine is situated near the town of Carletonville, 50 km south of Johannesburg, in the Gauteng Province. The 4 Shaft East complex comprises of the 4 Main Shaft and the 4 Subvertical Shaft. The main shaft is 9150 m in diameter and was sunk in 1984 to a depth of 2000.5 m. The 4 Subvertical Shaft is 7320 m in diameter and was sunk during 1985 to a depth of 3006 m. Full production was, however, only reached during 1988. Intensive mining is currently being done between levels 28 to 42. In 2002 a decision was made to start mining the shaft pillar, situated between 32 and 33 level in the subvertical shaft. Intensive studies showed that with selective mining scheduling the total mineÆs life could be lengthened by nine years if accessed timeously. The higher grades present in the shaft pillar can be mixed with lower grades from the same reef band presently mined 1500 to 2000 m away from the subshaft pillar area. The shaft pillar comprises 50 per cent of the remaining ore resources available. However, no provisions were made in the initial phases of the shaft development to extract the shaft pillar. Grades between 4300 cmgt to 2500 cmgt relating to 23 and 14 g/t are present in the shaft pillar and an estimated 50 tonnes of gold can be produced from this pillar in the period from 2007 to 2015. The pillar extraction is the key to DriefonteinÆs long-term gold profile. Every geological feature possible is represented in this to-be-mined pillar area. The pillar around the shaft is a complex structure featuring hard and soft lava, sedimentary structures/plane parting that moves against each other, quartzite capping, etc. Replacement of services and secondary support systems were designed, modelled and put in place. The original steelwork was not designed to cope with the various stresses and movements, which are expected once the extraction process is started. Existing shaft steelwork was used as far as practically possible. The principal behind the design was that the steelwork would be totally disconnected from the shaft lining over its total length of 132 m in the area of the reef intersection. This will allow the steelwork to move independently from the geological structures surrounding it, allowing a certain degree of distortion of the shaft lining, moderate twist and tilt to take place without interfering with hoisting operations.

Jan 1, 2005

By M Dixon, T OÆShannessy

Melbourne Water (MW) is responsible for a section of existing gravity trunk sewer servicing the Melbourne central business district, Docklands, Port Melbourne and other adjacent areas, known as the Melbourne Main Sewer (MMS). MW has determined that the 2.2 km long 110 year old sewer will not adequately service these areas into the future or provide the level of reliability needed. The Melbourne Main Sewer Replacement project (MMSR) involves the replacement of the MMS with a new gravity sewer main and six key manhole shafts, known as the Melbourne Main Reliever (MMR).The majority of the MMR sewer tunnel is situated south of the Yarra River and will be constructed using an earth pressure balance (EPB) tunnel boring machine (TBM) and will be lined with precast concrete segments. The tunnel will be 2.4 m internal diameter and will have a 1.8 m diameter glass fibre reinforced plastic (GRP) sewer pipe positioned within it. This project represents the first large-scale, modern, mechanised tunnel project in MelbourneÆs Yarra Delta Quaternary sediments. There is little precedence known to the authors of similar scale tunnelling projects in these particular geological units. During the recent detailed investigation and design stages of the project, significant effort has been spent on characterising the geotechnical conditions along the alignment and predicting the effects of the new tunnel on the existing environment. Some of the key design issues that have been identified and managed through the design process include short-term surface settlement due to tunnel excavation and the predicted effects on surface infrastructure, potential spalling of the steel fibre reinforced concrete segments under full TBM ram loads and tunnel lining design in geological units with highly variable stiffness.Construction of the new MMR sewer tunnel is expected to commence in mid-2008 with a capital cost of approximately $160 million. Sewer system commissioning is scheduled for completion by early 2012.

Jan 1, 2008

By R Smart

This paper examines the longer-term predictive capacity of the GAMSPATH computer program (developed by the Alberta Research Council) for a waste rock (sample MS) obtained from the Kaltim Prima Coal (KPC) mine, Kalimantan that contained significant framboidal pyrite. The model was refined by comparison with the net acid generation (NAG) and column leach tests to provide a basis for calibration of the model for both short term and longer-term reaction trends. The oxidation of pyrite during the kinetic NAG test using H2O2 provides oxygen directly at the mineral surface, generating a measured oxidation rate for pyrite of 1.16 + 10-6 mol m-2 s-1, whereas the oxygen diffusion limited rate in the leach column air equilibrated reaction was of the order 10-8 mol m-2 s-1 as calculated from the GAMSPATH thermodynamic database. Results indicated that the oxygen derived from hydrogen peroxide had to be included as a reactant (eg a reacting mineral) in the simulation to account for the direct availability of oxygen at the sulfide mineral surfaces. Optical investigations identified the proportion, particle size, and surface area, of framboidal and euhedral pyrite in sample MS. The reactive surface areas of the two morphologies were very different (eg ratio: 5:1). To manage these two different pyrite morphologies the GAMSPATH thermodynamic database was updated to include framboidal pyrite, with similar thermodynamic properties and oxidation rate to euhedral pyrite but with a particle size based on the size of the individual internal framboid microcrysts. This modification, with framboidal pyrite reaction based on the size of its microcrysts produced results that correlated well with the kinetic NAG and column leach test results when the measured (kinetic NAG test rate) and calculated (column test rate) pyrite oxidation rates were employed. This provided some confidence that the model could be extrapolated to indicate longer-term geochemical reaction trends.

Jan 1, 2003

By W Schauenburg

The increasing problems associated with the disposal of waste and the potential for use of by-products as underground filling material is a subject which is relevant not only to the coal-mining industry. The progressive scarcity of surface disposal facilities for by-products and waste has now become a problem in all sectors - not only for the mining industry. The considerable capacity offered by underground cavities which exists in various mining branches present an extremely promising potential in this context. The backfill process for coal mining developed by DMT in co-operation with Ruhrkohle AG make use of the cavities which are offered by the extraction of coal. It consists of several stages: The fine-grained by-products, such as sludge, filter cakes, fine refuse, ash and flue-gas cleaning residues are processed on the basis of specified formulations in metering, blending and pump facilities on the surface to produce pumpable pastes with precisely defined flow characteristics. On Walsum Colliery of Ruhrkohle AG up to 200 m3/h stowing paste is conveyed into the mine in pipes with diameter of 200 mm (8"). The delivery pressure of the pump operating on surface in conjunction with the hydrostatic pressure of the shaft pipe - amounting in total to as much as 25 MPa - makes it possible to convey the stowing material for a distance 12000 m underground up to the workings without any need for additional booster pumps. Via a pipe installed in the face, the paste is pressurised into the cavities in the specially fragmented rock in the mined-out zone well behind the support. Consistent use of stowage of caved-in goafs makes it possible to reduce surface settlement above the workings. Fires in the stowed rock are either smothered or completely prevented thanks to the exceptional air-tightness in the caved goaf and the cooling effects of the slurry stowage. The reduced danger of fire, the preventation of air-leakage and the positive effects on gas emission from the coal improve safety in the workings.

Jan 1, 1998

By R Piper

The Yabulu Nickel Refinery commissioned in 1974, processes lateritic ore via a modified Caron process to produce high quality nickel and cobalt products. The Greenvale orebody provided the sole source of feed to the plant until 1990 and with well-blended mineralogy, it resulted in stable and predictable plant performance, allowing the operating focus to be on reducing energy consumption and increasing ore throughput. This approach provided better economics than pursuing increased reduction efficiency. From 1990, increasing quantities of lateritic ore were imported from Indonesia, Philippines and New Caledonia. This introduced considerable variation in ore blends and instability in plant operation but also provided the opportunity to treat higher grade ore blends which yielded higher reduction efficiencies. Ore reduction is conducted using 12 Herreshoff multiple hearth reduction furnaces, using heavy fuel oil as reductant. The æROASTERÆ Project was approved in November 2000 with the primary objective of improving Nickel reduction efficiency by: Stabilising Roaster Operation via a structured program including focused maintenance efforts, capital equipment upgrades, new process measurements, process re-design and improvements to operating skills and process troubleshooting. A structured technical program including laboratory and plant studies using one plant furnace as a pilot unit to develop improved knowledge of how the reduction process proceeds in the Herreshoff furnace. Armed with this knowledge, optimised process conditions and a reduction process model were developed contributing to a significant demonstrated reduction efficiency improvement. Technical input to the project was provided by QNI with Hatch providing the engineering and project management for the $8.5 M capital work. Execution of the work included extensive consultation with the plant operators and close cooperation between the Hatch project team and QNI operations personnel ensuring installations were completed with minimal production disruption. A strong focus on safety management resulted in a number of brownfields project relevant safety initiatives being implemented resulting in project completion without a serious injury.

Jan 1, 2002

By S Basu, P M. Midhun

Electro-Slag Remelting (ESR) is a process capable of producing ingots with much lower centre-line segregation and better internal soundness than conventional ingot casting processes, as well as lower inclusion and sulfur concentrations. This process is commonly used for alloyed steels intended for critical applications like components of marine turbines, supercritical powerplants and rocket motor casings. It involves melting a metal electrode using electrical resistive heating in a slag pool and re-solidification of the resulting metal droplets beneath the slag layer, forming an ingot. Conventional ESR fluxes consist primarily of CaF2, Al2O3 and CaO, along with minor concentrations of SiO2, MgO etc. The emission of fluoride vapour from the molten slag causes several environmental and health problems, triggering a need to develop and reduce the fluoride content in the ESR flux. The present study aims to develop low-CaF2 alternatives to commercial ESR fluxes without compromising on productivity and the surface quality of the refined ingot. Thus, the modification in chemical composition needs to satisfy properties like electrical resistivity, viscosity and liquidus temperature, essential for efficient ESR operation. Since electrical resistivity and viscosity are highly dependent on the ionic structure of the molten slag, it is important to understand the effect of compositional modification on the structure of the flux. The present study employs Fourier-transform infrared spectroscopy (FT-IR) and Raman spectroscopy to investigate the structural variations in the modified fluxes, especially the effect of changing SiO2 concentration, in comparison with conventional CaF2-rich fluxes. Additionally, thermal analysis (Differential Thermal Analysis (DTA) and Thermogravimetric Analysis (TGA)) has been carried out to characterise the melting behaviour of the fluxes. The findings are compared with simulations carried out using the FactSage™ software, ver 8.3 (GTT Technologies, Aachen, Germany). Simultaneous analysis of liquidus temperature and the structure is expected to help in identifying the practical suitability of modified ESR flux compositions for industrial implementation.

Jun 19, 2024

By Heyer T, Wasmuth H-D

The concentrator plant of Fahrica Mine, that started production in 1977, was using nine Jones wet high intensity magnetic separators (WHIMS) of the standard type DP 317 with a 1.5 mm gap width for the production of pelletfeed concentrate. These WHIMS units were operating in parallel with a feed rate of 110 t per hour, with about 47 per cent Fe in the feed, and with a maximum of only 0.8 per cent magnetite. The field strength was about 1.2 Tesla. From 1986, the mine has adapted its production facilities to mine and concentrate larger amounts of lowgrade itabintes to maximise the use of the available iron ore reserves. An increase of up to three per cent magnetite in the magnetic separator feed, resulted in a reduction of capacity. This was due to partial clogging of the separation boxes at high field intensities, which were required for the recovery of the very fine hematite. Even by lowering the feed rate and the field intensity the problem could not be resolved. After extensive testwork in a pilot plant, with a Jones separator model DI' 90, the commercial WHIMS plant was modified into a two stage process using the existing nine Jones separators. The gap of the first stage with four machines was increased to a width of 3.5 mm, corresponding to a maximum field intensity of 0.7 Tesla. This first 'medium intensity scalping stage' is followed by a second 'real' high intensity WHIMS step with five machines operating at the original settings with 1.5 mm gap width and 1.0 Testa maximum field intensity. Feed rates of up to 320 tons per hour in the first and up to 240 tons per hour in the second stage could be achieved. The design concept of the concentrator at Fabrica mine is described using the old and new flowsheets. Details of the pilot plant testwork as well as actual performance data of the new plant section are presented.

Jan 1, 1993

Over the last 20 years the mining industry has played a key role in the further development of many Central Queensland and New South Wales communities. The number of people who are employed by the coal mining industry, the number of transient residents, shift work and the town facilities have all contributed to how the community views the coal mining industry. As the coal mining boom continues and the mining communities continue to grow more emphasis is needed to address the community concerns, specifically the socio-economic and environmental impacts that the coal mining companies may have on these communities. Central Queensland coal mining operations are situated in the Bowen Basin. Many of the towns in the region have been developed specifically for the coal mining industry with 13 - 50 per cent (Australian Bureau of Statistics, 2006) of people living in the communities being directly employed by the mines. The introduction of shift work, differing rosters and the mining boom have resulted in considerable changes to these communities with more transient contractors and permanent employees taking up residence outside of these communities. This change, however, has not been as noticeable in the Hunter Valley region. The Hunter Valley, situated in the Sydney Basin has two main mining towns, Singleton and Muswellbrook. Of the 37 173 (Australian Bureau of Statistics, 2006) people that reside in the communities 13.8 to 16.5 per cent (Australian Bureau of Statistics, 2006) are employed directly by the mines. The Hunter Valley is a large tourist destination due to its many vineyards; this results in a diverse community where mining is not the sole means of employment. Similar to Central Queensland the mining boom has brought more people to the Hunter Valley region, although greater percentages have taken up residence in the communities. The views vary between the two regions on the coal mining industry, specifically on the socio-economic and environmental impacts that the industry may have on the surrounding communities.

Jan 1, 2008

By S Prakash, P S. R Reddy, S K. Biswal, V N

The beneficiation potentials of two Indian iron ore slime samples by flotation technique using both cationic and anionic collectors are investigated. The direct and reverse flotation of slimes using oleic acid and dodecylamine as the collector have been studied using both conventional cell and flotation column to upgrade the iron content. The characterisation studies of iron ore slimes indicated that the samples are extremely fine in nature and contain substantial amounts of martite, magnetite, goethite, limonite, kaolinite and quartz beside hematite as the dominant iron mineral phase. The samples on average contain around 52 - 58 per cent Fe, four to seven per cent SiO2 and five to six per cent Al2O3. Taking into account the physico-chemical properties of the slimes, it is attributed that both the samples need to be deslimed by hydrocyclone before flotation. The results of both cell and column studies on sample I showed that a feed containing 57.5 per cent Fe could be upgraded to 65.5 per cent Fe at 50 per cent overall recovery with simultaneous decrease in SiO2 and Al2O3 content to 2.3 per cent and 1.8 per cent respectively by using oleic acid as the collector. In the case of sample II, it has been established that reverse flotation of silica is the better technique to improve both the grade and recovery of iron. Batch flotation studies in cell indicated that a product with 63.1 per cent Fe could be obtained at 40.0 per cent yield with 63.6 per cent iron recovery by the combination of reagents such as gum acacia, sodium hexa meta phosphate and dodecylamine. The performance of column flotation techniques is far superior to that of the cell with respect to both grade and recovery. The concentrates are found to be quite suitable for commercial exploitation as pellet feed.

Jan 1, 2005

By H M. A Ilyas, M Yahyaei, S Amari, M Darestani

Tailings are generally deposited on or near to the mine sites, occupying the site area, and being an economic burden for mining companies. Tailings can cause several environmental impacts, such as leaking toxic substances and contaminating underground water, emitting greenhouse gases, and damaging natural bodies and wildlife. Geopolymerisation of mine waste effectively transforms mine waste into more valuable, environmentally friendly products. Geopolymers are inorganic aluminosilicate materials that are widely applied in the construction industry (Araya, Kraslawski and Cisternas, 2020). The high performance and the low-CO2 emission of geopolymers make this type of material potentially more desirable than ordinary cement, given that manufacturing one ton of traditional cement generates approximately 0.8 tons of CO2. This process is seen as a promising method for repurposing tailings and reducing the environmental impact of mining waste. However, the environmental impact of geopolymer production is not sufficiently comprehended (Mazzinghy et al, 2022). Moreover, geopolymerisation of mine waste can offer several economic benefits, including reduced waste management costs (Amari et al, 2019; He et al, 2022), reduced raw material costs, and potentially lower production costs. However, there are several challenges that need to be addressed to ensure the economic feasibility of the process. This study aims to demonstrate the potential environmental impacts of geopolymer concrete produced using mine wastes and industrial by-products such as blast furnace slag (BFS). For this, a comparative cradle-to-gate life cycle assessment (LCA) methodology was conducted on different geopolymer mix designs, and the greenhouse gas (GHG) emissions results were compared to the ordinary cement. LCA can be used to evaluate the environmental impact of geopolymerisation of mine wastes in various ways. One important impact category is GHG gas emissions, which contribute to climate change and is the focus of this study. Another aim of this study is to evaluate the economic feasibility of the geopolymerisation of mine wastes via economic analysis (EA) to identify the key economic factors and trade-offs associated with repurposing mine wastes through geopolymerisation.

Aug 2, 2023

By S Karosiya, A Sharma, M Jain, N N. Viswanathan, M M. Pande, D Kumar, S Kumar

Desulfurisation of steel during ladle refining is governed by both thermodynamics and kinetics. For effective desulfurisation, thermodynamically, a low partial pressure of oxygen at the steel-slag interface and highly basic slag which possess a large sulfide capacity are desired. Kinetically, the mass transfer coefficient of sulfur in the liquid steel and in the slag determines the rate of desulfurisation. A fast desulfurisation rate is achieved by maintaining sufficiently strong argon stirring of the melt. In the present work, sulfur mass balance for steel and slag was performed after the desulfurisation treatment for several industrial heats. The decrease in the sulfur quantity of the liquid steel was more than the increase in the sulfur quantity of the slag. The discrepancy in the sulfur mass balance suggests that the sulfur could be transferred to: (a) the gas (via slag) phase; or to (b) the refractory (crucible) phase. The preliminary investigation of the refractory samples taken at slag-refractory and steel-refractory interface showed no interaction of sulfur with the refractory. Therefore, the possibility of sulfur transfer to the gas phase was verified by carrying out melting experiments of sulfur containing slag in an argon/air atmosphere. Sulfur containing slag was initially melted in an MgO-C crucible under an argon atmosphere followed by creating the air atmosphere just above the melt. Careful sampling and analysis from the cold spots of the crucible, much above the melt surface, revealed formation of a fibrous structure. Scanning electron microscopy energy dispersive X-ray spectroscopy (SEM-EDS) analysis of these fibrous deposits confirmed the presence of oxides and sulfides. Oxidation of sulfur containing slag under air atmosphere was also performed using a thermogravimetric analysis (TGA) set-up. It showed a constant sulfur loss in the range of 1300– 1500°C. Thermodynamic calculations carried out using FactSage™ ver 8.3 for these experimental conditions are in good agreement with the observed results.

Aug 21, 2024

A method is presented for calculating the number of cycles of stress to failure of a drill rod from operating variables; a mathematical analysis by Cooper of the stresses present in drill steels is applied to the forged collar, i1i-in. hexagonal hollow drill steel in common use in Australia; and an estimation is made of the dry fatigue limits at various points along a drill rod.It is shown that these calculations predict failure at a 'point where failure rarely occurs, and do not predict that failure should occur where it does in practice. The reasons for this anomaly lie in the neglect of bending stresses in the stress calculations, and in failuredue to corrosion fatigue at a preferred section rather than to simpler fatigue under non-corrosive conditions.INTRODUCTIONIt has long been recognized that failure of metal sections by fatigue is influenced both by the number of cycles of stress imparted to the section (N) and by the magnitude of this cyclic stress (8). In this paper, both Nand Shave been calculated for rods which have failed in service, and whose operating history is known. While the calculation of the number of cycles to failure is·simple, the calculation of stress is more complex. The results of stress caIculations predict that failure should occur only hi short rods, and then preferentially at the tip end rather than the shank end.However, in actual fatigue failures of steel rock drill rods, approximately 75 per cent. of such failures are close to the shank end, and failures near the tip end are rare.The usual explanation given for this high incidence of shank-end failures is that at the point of failure there exists, in forged collar steels of the type common in Australia "hardness valley" or. "m'etallurgical notch". This narrow valley in hardness is a weakness, and failure is expected to occur at the weakest point at the bottom of...

Jan 1, 1955

Now that Japan has achieved a remarkable high growth and joined the ranks of economic powers, she is required to take a stance proper to her new position toward international co- operation among Western advanced countries, economic/technical aids to Southeast Asian and Pacific LDCs, and others. As a field subject to such international cooperation programs, the development/use of coal resources has been winning increasing attention in recent years. Coal is positioned as one of promising alternative energies. With Southeast Asia plus Japan and the EEC serving as the two nucleus polar regions, total coal demand in the free world is projected to jump 37% in the year 2000 from the present level and steaming coal imports in the world, primarily destined for power generation, almost double the 1983 level. Such being the situations, Japan's tech- nologies related to coal resource development, if successfully introduced into coal-producing countries to help them boost their production, can directly lead to international cooperation that Japan is now required to fulfill. At home, however, the Japanese coal industry has been just on the wane due to various press- ing problems, such as wider differentials between overseas and domestic coal and deeper/ remorter mining. The industry was named atop in a list of the industries requiring restruct- uring at a meeting of the Economic Structure Study Committee held prior to the decision- making of the 8th coal policy, which slashed coal production planned for the last fiscal year of its life to some 10 million tons, a drop of as much as 50% from nominal output planned in the 7th policy. Under such circumstances at home and abroad, a pioneer study has been under way since fiscal 1984 as part of Japan's energy cooperation in the Pacific. The study is based on the Pacific coal flow concept hammered out by the Agency of Natural Resources and Energy of the MITI.

Jan 1, 1987

ON milling plants, where a large number of slime samples are collected each day, considerable work is involved preparing them for the assayer. The usual practice (as the material usually passes a 150-mesh screen) is to place the sample on a 30-mesh screen an 1 brush it through on to a cloth below. The sample on the cloth is mixed and a portion taken out for the assayer. A simple apparatus that would rapidly disintegrate and prepare these slime samples for the assayer was needed. The manufacturers of assayers' supplies appear to have overlooked apparatus for this class of work. To overcome the tedious preparation of slime samples by the above method, the writer devised the apparatus shown in the accompanying drawing (Fig. 1). It will be seen that the apparatus is a combination of a modified household flour sifter and a riffle sampler.The apparatus is made up in two parts, the top half consisting of a screen box carrying the disintegrator, and the bottom half a, riffle divider, the joints between the two being a slip-in fit. A lid is provided for the screen box. The disintegrator consists of an open drum, to which are soldered three rows of spring steel fingers. Each row has a number of steel fingers which, on rotation, press against the semi-circular screen in the screen box, forcing the slime particles through.To operate the apparatus an 8-in. sample tin is placed under each spout of the divider. The sample of slime, which is about 2 lb. in weight and usually in lumps, is placed in the screen box, the cover replaced, and the drum rotated about half-a-dozen times by means of the handle.The disintegrated sample is divided by the rifle divider into two portions, which are found in the tins under the spouts of the apparatus. One of these samples is taken, mixed on a sheet, and sent to the assayer.

Jan 1, 1922

The Hemersley Ranges are located in the North West Division of Western Australia and extend, approximately, from latitudes 21¦ 30! to 23¦S and longitudes 117¦ to 120¦E. Their mean height above sea level is'. 2,000 feet, and they rise from 600 to 800 feet. above the gen- eral level of the tableland which extends south-east from the coastal town of Roebourne.. Mt. Bruce, the highest peak, is 4,024 feet above sea level. The ranges are intersected by numerous north-flowing streams, dry during the greater part of the year, which have carved deep gorges, and it is in these that the principal deposits of orocidolite outcrop. The presence -of the mineral was first recorded in 1917., but the deposits remained untouched until 1937 when reports of a strong market demand for-croeido lice resulted in considerable prospecting and mining activity, particularly in Yampire, Wittenoom and Daleis Gorges. . It was established also that the blue asbestos belt extended from Willi- Wolli Springs north-west to Millstream Station, i.e.,-over a.:length of 180 miles and a width of 20 to 30 miles. A number of attempts were made to develop aid treat the Yampire and Wittenoom'Gorge deposits from 1938 onwards, but, for a ,variety of reasons, none of these-proved successful. Current production,. is wholly by Australian Blue Asbestos Pty. Limited at Wittenoom Gorge. This company commenced operations at the Wittenoom Mine in 1946, with an output of 601.8 long tons of- fibre,. and later transferred tb the Colonial Mine about one mile to the north-west. Output has been increased steadily to an' average of 12,000 to 14.,000.. long tons; per annum over the past six years. The total Western Australian output to the end of 1962 is reported by the Mines Department as 110,013 long tons valued at úAl2,403,728..

Jan 1, 1964

The Callie Underground Mine is located in the Tanami Desert 560 km Northwest of Alice Springs in the Northern Territory and is currently operated by Newmont Australia Limited with all mining activities carried out by Henry Walker Eltin. Callie was discovered in the region in 1991 with the box cut portal commenced south of the pit in 1995 and underground ore production commenced in 1998, after completion of the profitable Callie Open Pit. Development is currently occurring at depths of over 700 m with the decline currently below the 640 Reduced Level (RL) level (surface RL of 1400) with production haulage from 600 m below the surface. The combination of auto compression, a moderate geothermal gradient of 2¦C/100 m and diesel heat transfer causes the wet bulb temperature of intake air at the bottom of the mine to be typically 5¦C higher than on the surface. Wet Bulb temperatures can be up 5¦C higher again in active areas under secondary ventilation. During the 2002/2003 wet season many activities in the lower levels of the mine were stopped due to miners experiencing adverse heat related symptoms or the temperatures exceeding the Underground Heat Management Procedure stop job criteria of 32¦C wet bulb. To improve the working environment, combat the heat related illness and improve operational efficiency, a temporary 1.8 MW refrigerated cooling plant consisting of three 600 kW modules were installed underground to provide bulk air cooling for the lower levels of the mine during the 2003/2004 wet season. Several studies have been conducted in relation to future ventilation demands at Callie and all came to the conclusion, that some mine cooling will be necessary for Callie to meet the high production and development demands. This paper will describe the ventilation network at the Callie Underground operation focusing on the installation, operation and effectiveness of the temporary underground refrigeration plant.

Jan 1, 2005

At the outset I want to thank you for inviting a representative of the Chamber of Mines of South Africa to address you on the subject of mining and its future at this, the 13th Congress of the Council of Mining and Metallurgical Institutions. It is indeed a pleasure and an honour to be here. As many of you will recall, four years ago I, as the then President of the South African Institute of Mining and Metallurgy, had the privilege of hosting and welcoming you to the 12th Congress, the third such congress hosted by South Africa. The late Sir Ernest Oppenheimer, President of the 1930 Congress - the first to be held in South Africa - in his opening address foresaw the mining industry as "the most potent factor in spreading the benefits of civilisation over the sub-continent." We have come a long way since those halcyon days but, as I noted four years ago, it still remains true that mines have been, and will continue to be, the missionaries of development in many, if not most, of the Third World countries. I have been asked to address you on the subject of "mining" and to let my imagination roam widely so that delegates can leave the congress with a broader vision than they brought on arrival. It is a tall order and in my attempt to fulfil it I ask you to cast your minds back 120 years to a sultry, rustic day when the first diamond was discovered in South Africa at the now famous town of Kimberley. At that time the South African economy, predominantly pastoral, was still at subsistence level. Almost overnight the ensuing diamond rush turned the then British- owned northern Cape Colony's semi-desert scene into a hive of mining activity, every prospector bent on making his instant fortune. It was very little different, I am sure, to the gold rushes that hit California in 18+8, Australia in 1851, New Zealand in the 1860s. and the Yukon in the 1890s.

Jan 1, 1986

By H Holmstr÷m, A Widerlund, M Lindvall

At the Kristineberg mine in northern Sweden sulfide-rich, poorly buffered tailings have been deposited in five impoundments located along a valley. The tailings area has been remediated by applying a combination of till cover and flooding. A new method that has been applied to parts of the tailings area is to raise the groundwater table to saturate the tailings. By increasing pH by liming, thereby reducing metal mobility, a water-covered downstream impoundment is functioning as a trap for metals released from the till-covered impoundments upstream. In this paper the efficiency of the system approach used at Kristineberg is evaluated. Progressive changes of water quality from groundwater in the upstream till-covered impoundments via surface streams within the pond system to the water in the flooded downstream pond and its outlet are described. The results show that the system approach used in the design of the remediation programme of the tailings works as planned. The sulfide oxidation rate has decreased in the till-covered impoundments, and, as predicted, there is an out-flush of old weathering products which will continue for at least five years before equilibrium is reached. The groundwater in the till-covered impoundments has high concentrations of Fe, S, Mg and Zn, but the concentrations of metals such as Cu, Cd and Pb are much lower than before the till cover was applied. The limed pond is an efficient metal trap, with concentrations as high as 2.76 per cent Zn and 0.28 per cent Cu in the calcite-gypsum sludge developed on the tailings. The outlet from the flooded pond, and thus from the whole mining area, releases small amounts of metals to the recipient in comparison to the amounts that are in movement within the mining area. During the period May to November, 13 kg dissolved Zn and 111 kg particulate Zn was released to the recipient, whereas a total of 27 000 kg was transported by surface water within the impoundment system during the same period. The results from Kristineberg show that when planning remediation of complex mining areas, it is useful to adopt a system approach in which different methods are used on different parts of the area along the flow path of the drainage water, whereas flooding of the downstream pond is the key to the performance of the entire system.

Jan 1, 2003