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By M Gwynne

McArthur River Mine is located approximately 50 km south-east of the NT township of Borroloola and approximately 120 km from the Gulf of Carpentaria. The mine hosts a series of layered shallow dipping tabular orebodies, with current reserves of approximately 22 Mt, grading 13.6 per cent zinc, 6.3 per cent lead and 65 g/t silver. Owing to economic considerations, the lower most orebody (No 2 orebody) has been selected for initial mine production. Following the completion of the No 2 orebody, mining will transfer to the lower grade and metallurgically more complex No 4 orebody. The mine employs a total of 235 people throughout the underground and surface areas. The majority of the workforce work a fly-in/fly-out, seven-day- on/seven- day- off roster with 12-hour shifts. Presently, a further 18 contractors are involved in trial mining areas. An apparent dip room and pillar technique has been selected as the primary mining method. Officially opened in September 1995, the current production rate of 1.2 Mt per annum will eventually escalate to 1.5 Mt per annum. Though most production is achieved through the use of conventional loaders, trucks and jumbos, the recent development of methods utilising scrapers, hand held machines and modified jumbos has produced encouraging results in steeper dipping areas of the mine. Initial production has been hindered by several factors including extensive and largely unpredictable faulting, worse than expected ground conditions and difficulties with some mining equipment. Compounded by the relatively thin 3 m ore seam, initial mining resulted in high dilution and poor production. This paper describes the development of unique planning and production systems that were evolved in order to overcome these problems and achieve the targeted production rates and head grade.

Jan 1, 1998

By N Green

The production of approximately 550 000 tonnes of æwhiteÆ or chrysotile asbestos fibre between 1972 and 1983 at the Woodsreef Mine in New South Wales left behind 24 million tonnes of tailings and 75 million tonnes of waste rock and overburden. The Woodsreef Magnesium Project is designed to extract the 23 per cent magnesium content of the tailings to produce magnesium metal and magnesium alloys, the latter with aluminium from the New South Wales smelters. Further value will be added to the tailings waste with the downstream production of magnesium alloy sheets and cast products for the transportation, electronics and other consumer goods industries. The mine opened in 1918 and closed in 1983 becoming derelict. There has been minimal expenditure on rehabilitation compared with that recommended in reports commissioned by the New South Wales Government. The tailings dump has been designated æasbestos wasteÆ by the EPA and will be eliminated through the magnesium refinery process of the Project with the dissolution of the remaining chrysotile fibres. The benign waste from the refinery process will consist principally of silica with some iron. Significant energy savings are expected from innovative new processing technology, in the alloy formation and in the application of the light weight, cast and sheet metal products, particularly in the transportation industries. Globally, there are very large resources of similar tailings with hundreds of years of useful productive lives that could be converted into the metal of the 21st Century, magnesium, and its alloys, to assist pollution and greenhouse gas reduction from transportation industries. No new mining and minimal feedstock preparation is required. Unlike potential producers using carbonate rocks such as magnesite and dolomite that have to be mined, the asbestos tailings composed principally of non-fibrous serpentinite do not contain any carbon dioxide.

Jan 1, 2002

By T Warren, A Dempster, N Barnett, J Oh, S Saydam

The influence of the insulating properties of lunar regolith was shown in the sub-surface temperature tests performed by the Apollo missions, with no significant temperature fluctuations occurring below 80 cm from the lunar surface even if surface temperatures fluctuated >200 K between lunar days and nights (Heiken et al, 1991). Key to the thermal insulation of the regolith is its low density near the surface having the greatest insulative effect with the lunar regolith becomes denser with depth. As the regolith becomes denser, the thermal conductivity of the regolith increases (Hayne et al, 2017). Thermal modelling software, Oxford 3D Thermal Model, which mapped the lunar surface temperatures effect on the thermal influence at the lunar south pole sub-surface over a lunar year found areas with brief illumination with surface temperatures >120 K could have sub-surface temperatures <112 K that can sustain long-term stable water ice at depths of <30 cm (Paige et al, 1992; King et al, 2020). Further to this, the observation of water molecules or other hydroxyls by the NASA/DLR Stratospheric Observatory for Infrared Astronomy (SOFIA) in non-permanently shadowed regions (PSRs) on the Moon, indicated that water could be present beyond PSRs (Honniball et al, 2021). Using the Oxford 3D Thermal Model in selected regions in the lunar south pole, can provide insights into the location and depth of cold traps that could sustain water ice.

Nov 29, 2022

By A Golev, G Corder

The concept of the circular economy has been gaining traction both in Europe and China.In Europe the Ellen Macarthur Foundation see their mission as accelerating the transition from a linear take-make-dispose economic model to a circular model that is restorative and regenerative by design, and aims to keep products, components and materials at their highest utility and value at all times. The Chinese Circular Economy initiative is effectively a sustainable consumption and production program utilising cleaner production, industrial ecology and life cycle management approaches to meet the national challenges of maintaining rapid economic growth while simultaneously enhancing environmental quality and maintaining social progress. A key characteristic of both of these circular economy initiatives is to design out, reuse or minimise the generation of waste.The mining sector, by definition, is represented by linear rather than circular activities, by supplying resources to society. However, being one of the world’s largest waste generators, the sector can adopt similar logic to that of the circular economy to improve its sustainability performance. Waste is a critical issue along the whole metals value chain, from mining waste to eventual end-of-life products, such as scrap steel from construction and demolition waste or the growing problem of electronic waste.CITATION:Corder, G and Golev, A, 2016. Can mining be part of the circular economy or is it squaring the circle?, in Proceedings Life-of-Mine 2016 Conference , pp 121–123 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jun 28, 2016

The Golden Cross mine is situated in a sensitive environmental setting on the North Island of New Zealand, about 150 kilometres south-east of Auckland and about 8 kilometres NW of Waihi. The mine is situated in dipping topography, with volcanic ashes making up soils, fault zones running through the area, and rainfall of about 3m per year on average. Water management is the most demanding environmental challenge on site, and a comprehensive programme has been put in place to address tailings deposition, long term tailings storage, and water treatment during and after operations.

Jan 1, 1995

By S T. L Harrison, J-P Franzidis, M Becker, N Dyantyi, J L. Broadhurst

Gold mining is a major contributor to the South African economy, but is associated with several environmental issues such as acid rock drainage (ARD). ARD is due to the oxidation of sulfides in gold tailings and is characterised by low pH, high sulfate content and elevated concentrations of heavy metals that contaminates scarce water resources and subsequently limits the use of such waters without treatment. Treatment strategies can only be implemented if the acid-generating potential of mine wastes is effectively predicted using common static and kinetic acid prediction tests. The choice of test performed is usually a trade-off between the number of samples, speed of the test, cost of the analysis and level of information obtained. Often the results of these various tests are not consistent with one another. This study uses mineralogy to analyse biokinetic test results of a Witwatersrand gold sample in order to improve the understanding of behaviour of mine wastes under different ARD prediction test conditions.All ARD prediction testing methods investigated in this study showed that the gold sample is considered to be potentially acid forming with net acid generation (NAG) pH of 2.5 and net acid producing potential (NAPP) of 36 kg H2SO4/t. Results of biokinetic test showed that ANC was consumed within four days confirming acid producing potential of the sample with low neutralising potential as observed from acid base accounting (ABA) (4.7 kg H2SO4/t) and mineralogy (4.9 kg H2SO4/t). Results generated in this study demonstrated importance of mineralogy in both understanding the controlling reaction mechanisms playing key role in ARD formation and its capability as a prediction technique. Future work will focus on the application of mineralogy in interpreting the behaviour of mine waste under different static test conditions with a view to assessing their applicability and reliability in estimating potential acid generating characteristics of gold tailings.CITATION:Dyantyi, N, Becker, M, Broadhurst, J L, Harrison, S T L and Franzidis, J-P, 2013. Use of mineralogy to interpret laboratory-scale acid rock drainage prediction tests - a gold case study, in Proceedings World Gold 2013, pp 519-526 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 26, 2013

By Laros T, Schoenbrunn F

Deep paste thickeners have been proven capable of consistently producing medium to low slump material for disposal without the need for filters. However, proper design of the entire system is critical for a paste project to he successful. Some of the factors that must be carefully considered in design of a paste system include thickener sizing, tailings inipoundineni design, underflow pumping, and control system.

Jan 1, 1998

By J Player, R de Vries, E Villaescusa, R Hassell

Concerns regarding the uncertainty of stope dilution at MMG Dugald River Underground Mine led to a trial stoping program to acquire full-scale comprehensive geotechnical information and to test the validity of the geotechnical and mining parameters. Given that stope production drilling and blasting is a key parameter that influences dilution a detailed stope blast vibration project was implemented. The stoping trial consisted of 19 stopes, a combination of both sublevel open stoping and longitudinal and transverse stopes, of which five were monitored for blast vibrations.The blast vibration project utilised 20 triaxial geophones and five triaxial accelerometers installed in the hanging wall of five monitored stopes. For each stope fired, full blast waveforms were recorded along with detailed information of the blast design. Analysis of the blast vibrations used the Holmberg-Persson approach, with potential damage determined by the relative (Holmberg-Persson) K values.The project identified the benefit of parallel, breakthrough slot holes in reducing blast vibration damage. The results show that radial rings toeing or stabbing into the stope wall lead to higher vibrational energy yield which contributed to increased overbreak. Hanging wall overbreak, by comparison, with parallel, breakthrough blastholes with 1 m standoff, combined with lower density explosive charging (0.8 EP) and increased delays limited blast vibration damage. A number of ring burdens from 1.4 m to 2.5 m were monitored with no difference in terms of energy yield being identified. Changes to the stope drill and blast design are suggested to better control the large blast vibrations and recommendations from this project have been implemented into future mine design and planning.CITATION:Hassell, R, Villaescusa, E, de Vries, R and Player, J, 2015. Stope blast vibration analysis at Dugald River Underground Mine, in Proceedings 11th International Symposium on Rock Fragmentation by Blasting, pp 221–232 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 24, 2015

By B Petrie, D J. MacKenzie

The Macraes mine is producing 175 000 ounces Au/year from a mesothermal deposit with total resource of at least five million ounces. The deposit is hosted in greenschist facies Otago Schist, a widespread Mesozoic metasedimentary sequence that forms the basement of the South Island of New Zealand. Mineralisation was controlled by a late metamorphic shear zone which had both brittle and ductile structures accompanying mineralisation. The shear zone dips shallowly (about 15¦), subparallel to the shallow-dipping regional foliation, and mineralization occurred during southwesterly-directed thrust motion. Mineralisation occurred under greenschist facies conditions (about 350¦C) during uplift of the Otago Schist in the late Jurassic. No syn- or late-metamorphic igneous bodies occur within the schist belt. The hosting shear zone is up to 120 m thick, and is traceable for more than 30 km along strike. The shear zone is truncated by an underlying low angle normal fault that juxtaposes mineralised lower greenschist facies rocks on to unmineralised upper greenschist facies rocks.

Jan 1, 2004

By Ward S. D

When using grinding time for a batch grinding operation as a measure of the energy usefully employed for comminution; the assumptions made are often not valid. A kinetic approach, based on a study of the rate constant, or k, value for the material is more acceptable. The Rosin-Rammler distribution equation has been theoretically derived from the general equation for first-order batch grinding. This relationship, incorporating a term for k, was used in a study of the rod milling of a binary mixture of quartz and limestone.The presence and degree of interaction between the components was considered by comparing the size distribution of the mill product with that predicted by assuming the overall distribution equal to the sum ofthe distributions given by the Rosin-Rammler equations for each component weighted according to its volume fraction in the mixture. The interaction was found to be complex and particularly sensitive to particle size.A simple graphical procedure was devised to predict the size distribution of the mill. Product components and this gave a fairly good approximation to experimental data.

Jan 1, 1976

By D Trembath

"One of the key risks to the success of a mining operation is the accuracy of the estimates of both the economic mineral and any accompanying contaminants contained within the ore reserve. Misrepresenting the mineralogical characteristics of the ore reserve can result in considerable financial investment in an operation that is in fact uneconomic and/or carries a significant environmental risk. This in turn can result in an unplanned closure of the operation with associated financial, environmental and social legacy. One way to minimise this risk is to undertake high density resource sampling to gain a better understanding of the resource characteristics. However, such an approach is expensive, time-consuming and often impractical. The risk of not having an accurate model is typically assessed during a feasibility study utilising a Monte Carlo simulation method or some form of sensitivity analysis. Importantly, the error margins used in these types of analyses are often based on subjective or very general criteria such as ‘professional judgement’ or an ‘industry standard’. This presentation introduces a conditional simulation method for developing a range of mineralogical distribution models. Based broadly on the method outlined in Journel and Kyriakidis (2004) the proposed approach develops this work further. It combines a geostatistical analysis of the available data with automated mine design software to provide a rigorous method for assessing the confidence in the mineralogical characterisation of the ore reserve. The process is based on simulating possible mine plans based on the statistical characteristics of the available resource sample data. These plans are then assessed against high resolution resource models generated using a Sequential Gaussian Simulation method. The outcome is confidence intervals for important characteristics of the ore reserve estimate relative to the actual resource. These confidence intervals are an objective and transparent measure important for assessing closure risks and represent an improvement over common practice.CITATION:Trembath, D, 2016. A simulation approach for improving the assessment of closure risks, in Proceedings Life-of-Mine 2016 Conference , pp 36–38 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Jun 28, 2016

By D Matthews

With the increasing demand for quality-specific ores, coupled to the reality of even more complex minable iron ore reserves, there is a growing need for real-time online material analysis of plant feed and product. For a full system mass balance, the combination of conveyor weight scales, flow and density transmitters are required. This alone does not guarantee product grade. For optimal process control and material blending, real-time online elemental scans in combination with the mass weighting is becoming more than a ‘nice to have’. These elemental scans located at key positions, enable the mining and plant operators to make informed decisions timeously, therefore ensuring product outcomes and optimising operations where material grade and the control thereof are important. Typically, these units follow the mining and primary crushing circuit as well as on the product conveyors, low-grade conveyors and tailings lines. Deployed after the primary crushing circuit, an analyser acts twofold:1. guaranteeing mining production, and2. processing facility feed blends.The analysers on the product conveyors guarantee saleable product. The analysers on the low-grade conveyor and tailings line complete the elemental balance and identify the potential of the low-grade and tailings for future processing. The system losses (accumulation) are calculated to determine the amount of material ‘lost’ during the processing stage which does not report to any of the aforementioned and can be used as a measure of the efficiency of the plant. A further advantage of the real-time online analysis of material is that immediate intervention can be taken in the event that the product is not of saleable product grade requirements. In the Khumani Iron Ore mine, Northern Cape, South Africa, run-of-mine (ROM), product and control analysers were included. Run-of-mine and product stockpiles are certifi ed, low-grade material is not currently directly measured, but calculated from the internal control loops. Iron units are reconciled between product and feed material.

Jul 11, 2011

For many years the mining industry at Broken Hill, New South Wales, Australia, was carried on with just sufficient attention to underground ventilation and hygiene conditions as to keep pace with necessity. However, the campaign against industrial diseases in other mining centres (notably South Africa) some years ago, brought the subject more prominently before local mining engineers.Throughout a period of several years, gradual improvement has been effected, and at the present time industrial hygiene is considered an essential part of the production operations of the various companies.In June, 1918, a move was first made by the Board of Trade advising an inquiry respecting the health of employees engaged in the metalliferous mining industry in the State of New South Wales. At the end of that year, a recommendation was made to the State Government by the Board of Trade that a Techni.cal Commission of Inquiry be appointed to examine and report upon the dust conditions of the metalliferous mines at Broken Hill. Following this a second report presented in 1919 made specific recommendations in respect to the scope of such a Commission. A Technical Commission of Inquiry was accordingly appointed and arrived in Broken Hill to inquire into the prevalence of miner&apos;s phthisis and pneumonoconiosis in the mines. Early in March, 1920, the examination of all mine employees commenced, and the work was completed in June, 1920. The full report of the activities of the Commission and the conclusions arrived at was officially published at a later date. Based on the recommendations...

Jan 1, 1940

By J Millard, K Sexsmith

BHP Billiton Diamonds Inc (BHPB) has been mining kimberlite from several pipes at the Ekati Diamond MineTM located 300 km north of Yellowknife since 1997. Processing of kimberlite produces coarse kimberlite rejects (CKR, 0.5 to 8 mm) and fine tailings (less than 0.5 mm). The CKR are trucked to the Coarse Kimberlite Reject Storage Area (CKRSA) located within the drainage area of the tailings impoundment (Long Lake Containment Facility û LLCF). The coarse rejects are wet when dumped and are expected to freeze as a result of annual average site temperatures of about -10¦C. Seasonal thawing occurs along the exposed edges of the pile. Monitoring of the CKR indicates they have consistently low sulfur concentrations (0.4 per cent) accompanied by high acid neutralisation potentials (78 to 337 kg CaCO3/t). Carbonate concentrations vary from 0.9 per cent to 3.4 per cent. The CKR are therefore not potentially acid generating based on conventional laboratory testing. Monitoring of stagnant pools on the tundra downgradient of the CKRSA has shown that water originating from the pile resembles acid rock drainage (ARD). Water pHs can be less than 4, and sulfate concentrations are typically in the thousands of milligrams per litre. Conventional flowing toe seeps are pH neutral but also have elevated sulfate concentrations. Investigations conducted in the summer of 2002 revealed no evidence of in situ acid generation. Pore water chemistry is probably controlled by oxidation of fine-grained pyrite, neutralisation of acidity by calcium and magnesium carbonates and silicates, and freeze concentration of pore water. The acidity observed in tundra pools downgradient of the CKRSA appears to be a result of leaching and mobilisation of iron under reducing conditions at the CKR-soil contact. Downstream of the CKRSA, reduced iron oxidises and is hydrolysed on contact with the atmosphere causing pH depression. BHPB is applying the findings of this study by constructing facilities containing kimberlite on granite pads. These in turn are placed on frozen ground so that the foundations remain frozen and limits contact of wastes with the acidic soils.

Jan 1, 2003

By S Brunt

A new elution process has been developed for the recovery of gold thiosulfate from ion exchange resins. This is a synergistic system which utilises a base eluent, such as nitrate or trithionate, along with an additive that improves the elution of gold. The use of the additive enables lower eluent concentrations to be adopted, as well as weaker eluents such as chloride. With the chloride + additive system, complete gold elution can be achieved within 12 BV when utilising an eluent flow of 5 BV/h. When the eluent flow is reduced to 2 BV/h, complete elution is achieved within 8 BV. Similar results are obtained when using nitrate or trithionate as the base eluent. A copper pre-elution process has also been developed, whereby the majority of copper can be stripped within 5 BV prior to gold elution. This method is a significant improvement on the ammonia + oxygen pre-elution previously suggested. The development of a new analytical procedure based on ion chromatography with UV detection is also described for the quantification of polythionates in leach solutions and loaded on resins. This enables analysis to be carried out for resin-in- leach type processes where the polythionates are being formed in solution and simultaneously adsorbed onto the resin.

Jan 1, 2007

By A Scott, S Kanchibotla, J Segui

The æMine to MillÆ concept is not new. In fact, it is the underlying objective of every mine. Unfortunately, in many operations the management structure and reporting systems have created independent æbusinessesÆ within the overall mining enterprise. This has created incentives to optimise the performance of each component of the operation rather than the enterprise as a whole. The Mine to Mill Conference organised by The AusIMM in October 1998 documented a number of case studies where run of mine ore was not presented to the mill in the best possible form for subsequent stages of comminution and processing. A number of examples were quoted where modest expenditure in the mine had a very significant impact on downstream process efficiency and hence costs. Examples included:the minimisation of fines in the mining and handling of iron ore; the management of fines in blasting and crushing for a copper heap leach operation; the control of blast damage, mining recovery and dilution in open cut coal operations; and the optimisation of mine fragmentation to increase mill throughput in gold and copper operations. The latter topic has been trialed at a number of mines and is referred to as æMine to MillÆ fragmentation. A number of æMine to MillÆ type investigations have been reported in the literature including Pease et al, Simkus and Dance, Nielsen, Scott et al, Kojovic et al, Kanchibotla et al, Bulow et al and Hawkes from the 1998 Mine to Mill Conference. In addition, a number of studies are in progress or the results have yet to be published. These studies have demonstrated the feasibility of obtaining the data required to characterise the breakage behaviour of ores and to model the impact of changes to blasting practices on process efficiency. Characterising the breakage behaviour of ores is a vital step in the management of fragmentation and comminution in mines. The characteristics of importance to fragmentation fall outside the parameters conventionally required to describe the economic geology of a deposit. æMine to MillÆ optimisation of fragmentation means that the mining engineer needs information from the geologist in order to be able to help the metallurgist. Providing this information requires close interaction between these groups.

Jan 1, 2000

By V S. Vutukuri, K Shariar

This paper present a modification of selected strength criteria for coal and rocks containing weakness planes. The three strength criteria, namely, Murrell (1965), Bieniawski (1974a, 1983) and Hoek and Brown (1980) have been assessed for coal (containing weakness planes) by using triaxial strength tests. Triaxial test results indicated that the strength of coal specimens depends on the weakness plane inclination with respect to the principal stress direction. However it is shown that with increasing confining pressure, the effect of weakness planes inclination on the strength of coal specimens is reduced. The constant parameters in the three failure criteria for coal are determined for each weakness plane inclination by using the Datafit computer program (nonlinear regression analysis) and the results are compared with other rock types.

Jan 1, 1999

Experimentation with new methods, materials and machines to improve primary drilling and blasting is a continuing practice in the hard rock quarry operations of the BMI Group. These experiments will continue in the remote hope that one day we may produce a perfect feed for the primary crusher. As this day is still a long way off, the necessity to break oversize boulders will be a problem to be faced for the foreseeable future. To gauge the efficiency of primary blasting from this point of view, BMI monitor the tonnes of material crushed in the primary crusher for each boulder requiring secondary breaking. Over a period of about 10 years, this figure has improved from 30 tonnes througi the primary crusher per boulder, to 100 tonnes per boulder at our largest operation, which produces in excess of 2 million tonnes per year. Evan with this improvement, 25,000 boulders require secondary breaking each year This paper does not provide the solution to this problem, as like primary blasting, it will have to be constantly reviewed as new products and techniques become available. The paper sets out all of the methods tested or considered for test by BMI and our nnnn1iicinnq nn Chair PffPr.tivPness . Economic

Jan 1, 1978

By N DavT

Pyritic tailings from the Quirke mine waste management area in Elliot Lake, Ontario were subjected to long-term column leaching experiments in two PVC columns measuring 122 cm in height and 15.2 cm in diameter. At the end of the leaching experiments, the columns were taken apart and sampled at 5 cm intervals for quantitative mineralogical and geochemical characterisation. Each column is represented by a total of ten samples. Polished sections, prepared from the samples, were studied by optical microscopy, scanning electron microscopy, electron microprobe and X-ray diffraction techniques. Mineral quantities were determined based on linear mixing models of whole rock and mineral compositions. Pyrite particle sizes were measured by image analysis interfaced to an electron microprobe. The tailings are coarse-grained (ie 96 per cent passing 75 ¦m) and composed of quartz, K-feldspar, muscovite and pyrite. Pyrite is homogenous and occurs as liberated particles displaying angular to subangular boundaries with strong dissolution features at the end of the leaching experiments. Based on up to 3518 particles measured in a sample, pyrite particle sizes range from 1.7 to 78.6 ¦m. Pyrite concentrations vary from 0.1 to 7.8 wt per cent, depending upon their location from the top of the columns. These correspond to pyrite losses of up to 99 per cent from their original level of 7.2 wt per cent in the upper half of the columns. Pyrite oxidation rates in terms of moles of pyrite oxidised per mole of pyrite per square metre per second were calculated from the effluent data and the measured pyrite particle sizes. The rates vary from 1.71+10-12 to 5.78+10-10 mole+m-2+s-1 within the 1.5 to 3.5 pH range for 10 ¦m particles. These results are comparable with the published oxidation rates at pH 2 where the dissolved oxygen is the only oxidant.

Jan 1, 2003

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