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IN the ordinary course of treatment of pyritic ores or concentrates, whether for the recovery of copper, tin, gold, or any other contents, the removal of some or all of the sulphur contents is almost invariably the first step. The practice is to roast the ore in a current of air so that the sulphur is oxidized to sulphur dioxide and the metal constituents are changed from sulphides to oxides. If it is feasible to produce sulphuric acid in conjunction with the treatment process, the sulphur can be usefully employed; but if this is not possible, then the sulphur contents not only become a dead loss, but tend to be a nuisance as well in the surrounding neighbourhood.The problem of avoiding this disadvantage and making economic use of the sulphur has had attention from the early days of scientific metallurgy, and in the past few years very special efforts have been made both to drive part of the sulphur as such from pyrites and recover it, and to replace the sulphur in the S02 by carbon, with the production of carbon dioxide and precipitation of the sulphur. So far economic success does not appear to have crowned these efforts. In the case of pyrites, of the two atoms of sulphur which go to make the FeS2, the first is very lightly held. Dissociation is effected by heat alone at a temperature which is given by Peters as 5160 C.Now, sulphur and hydrogen will unite or burn together, with the production of sulphuretted hydrogen. It occurred to the writer that, if hydrogen was passed over pyrites heated to the above temperature, this first atom would be snatched away by the hydrogen, and could possibly be economically recovered.The first step consisted of some simple laboratory experiments. From these it was found that this first atom was readily removed, and that the extraction was practically complete. Further, the reaction took place at a low temperature, a dull red heat being sufficient.

Jan 1, 1923

A formalised approach to risk management has been regulated in the United Kingdom (UK) since the Construction (Design and Management) Regulations 1994 (CDM) came into force in 1995. These regulations have recently been revised in 2007. In response to a series of significant claims around the world the UK insurance industry published a code of practice for risk management of tunnels in 2003, and that was followed by an international version in 2006. Probably one of the main contributions of this new approach is the recognition that health and safety risks can best be addressed early in the process, initially by the owner and then by the designers. One of the aims is to achieve elimination of risks by good design. The advantage of a risk-based approach to tunnel design is that all risks are considered explicitly, and any residual risks in the design are communicated clearly to the construction team. This paper looks at the implementation of this formalised approach to risk management with reference to the A3 Hindhead Project. The design has been undertaken in accordance with the requirements of the CDM regulations, and also in compliance with the Joint Code of Practice for Risk Management of Tunnel Works in the UK. An outline of the risk-based design process used is provided, including examples.

Jan 1, 2008

Reliable sampling practices are the key to optimising the market value of a mining or chemical companyÆs stock. As a result, reliable sampling standards are needed. However, current standards on sampling present the æuniversalÆ Pierre GyÆs Sampling Theory in an ambiguous way. For example, the Total Sampling Error is divided into its basic components only in informative annexes, when it should clearly be the main body of the standard. Sampling standards are presently loaded with statistical concepts that actually belong in the annexes. Furthermore, many non-probabilistic sampling procedures are still mentioned which do not belong in a standard. The objective is not to criticise existing sampling standards, but to suggest better ones. Sections on the Sampling Theory should divide the Total Sampling Error into a sum of its nine basic components. Then, and only then, should they be addressed one at the time. Sections for establishing a sampling protocol should address the in situ Nugget Effect, the Fundamental Error, the Grouping and Segregation Error, the Long Range Quality Fluctuation Error, and the Periodic Quality Fluctuation Error. Sections for sampling methods, mechanical sampling, manual sampling, grab sampling, sampling of trucks, railroad cars, bags, and drums should be part of the critically important practical implementation of a recommended sampling protocol, called Sampling Correctness. It includes the increment Delimitation Error, the increment Extraction Error, the Weighting Error, and the Preparation Errors. Unless these errors are addressed one at a time, in an unambiguous way, the standard will remain ineffective in solving sampling problems for the industry. Therefore, a strategy is needed to create better standards required at the early stages of new projects, both by manufacturers of sampling equipment and by engineering firms. A logical sequence of important points that cannot be ignored in any sampling standard is presented. An analysis of the practical use of the words sampling uncertainty versus sampling error is recalled. Finally, as a practical example, an in-depth analysis of the correctness of cross-belt samplers is presented, and emphasis is placed on the correct ways to perform bias-tests for mechanical sampling systems.

Jan 1, 2005

XstrataÆs Mount Isa Copper Operations (MICO) is a world-class mining operation producing in excess of six million tonnes of copper ore per annum from its underground mining operations. The Mount Isa Copper Mines have been producing from the underground operations since 1952. Current copper reserves are 87 million tonnes at 2.8 per cent Cu, extending the mineÆs life beyond 2018. The underground orebodies in the copper mines are divided into nine distinct areas and are mined by sublevel open stoping techniques. One of the nine areas is the hanging wall lens, which consists of two sections; a northern lens (NHWL) and a southern lens (SHWL). The SHWL has a proven reserve of 2.2 million tonnes with an average grade of 2.9 per cent copper. Stopes in the SHWL have an average size of approximately 245 000 tonnes and will be extracted in a primary-secondary mining sequence, producing ore until 2009. This paper discusses various aspects of stope design applied in the southern hanging wall lens, including sequencing options and the rationale for the selected sequence. Design features are also discussed and how they have evolved as more is learnt about the SHWL orebody. These include mass firing techniques, focusing on minimising ground vibrations, using seed wave form analysis and prediction. Rock mechanics issues will also be discussed, in particular the reinforcement of stope hanging walls and the use of instrumentation to monitor behaviour within the hanging wall.

Jan 1, 2006

The key elements required to maintain final wall integrity and hence slope reliability are discussed. Failure modes and mechanisms in highly structured rock slopes are then reviewed. The impact of blast disturbance on rock masses in general is discussed and the relationships between the slope failure mode and blast disturbance is then explored. The question of how designers can incorporate such impacts into the slope design process is then addressed. This necessarily involves rock mass strength consideration and two case studies are used to demonstrate a recent dilemma in this regard. Some progress regarding this dilemma has also been reported for the second case study. The major conclusions are that: Slope reliability can be severely compromised if the slopes are damaged by blasting and this impact is not often incorporated into the slope design process. The location and extent of the blast damage will impact some slope failure geometries and failure mechanisms more than others. The concept of failure mechanism specific vulnerability needs to be further developed. The evidence indicates that this aspect is even more extreme for weak ground conditions. The quantification of the reduction in rock mass strength due to blast damage is not well understood and limitations exist with some recent guidelines. The use of the Hoek-Brown criterion for slopes in blocky rock masses requires considerable judgement. Further applied developments in this regard would be very beneficial to the industry. Some promising methods to incorporate blast damage estimates into the slope design process are emerging. It is good business to minimise final wall blast damage otherwise a whole range of issues relating to reduced slope reliability emerge.

Jan 1, 1999

By Seymour C. A

Peak Downs Mine, a five million tonnes a year coking coal operation, strip mines a, series of Permian coal seams. Seam thickness ranges from 0.5m to 5.0m, and seams consist of multiple coal plies with intermediate shale and mudstone partings. Four 47m3 draglines strip overburden to a cover depth of up to 60m. Coal loading is by 12m3 shovels and 14m3 to 18m3 front-end loaders. Parting removal takes place where the layer of shale and mudstone between seams exceeds 0.3m in thickness. Parting has been removed by a wide variety of earthmoving equipment in order to evaluate the problem and to select the optimum method from an operational and economic standpoint. Multi-seam stripping of the thicker seams is accomplished by working draglines from spoil- pile benches. Coal seams thinner than lm are recovered by dragline and stacked in the pit for later loading.

Jan 1, 1974

By B Goebel, G Baker

The Australian Magnesium Corporation Limited (AMC) is constructing the worldÆs largest magnesium metal plant at Stanwell, 24 km west of Rockhampton in Central Queensland. The $1.3 billion, 97 000 tonne per year plant is scheduled to commence production in mid-2004. It will produce magnesium from magnesite using the innovative, locally developed AM Process Technology which enables magnesium to be produced at internationally competitive costs. The technology will enable the plant to have the lowest environmental impact of all the available magnesium production processes. The project will meet current world-best-practice environmental and emission standards while on a life cycle analysis, magnesium produced by the Stanwell plant will result in net reduction in global greenhouse gas emissions. Magnesium is the lightest of the structural metals. Its high strength to weight ratio and ease of casting and machinability makes it particularly suitable for the manufacture of light weight automotive components. Reducing the weight of motor vehicles through the use of light weight materials such as magnesium is the current primary means adopted by the automobile manufacturing industry to improve motor vehicle fuel efficiency and the lowering of emissions of greenhouse gases. The Australian Magnesium or AM process has been developed over the past decade by AMC in conjunction with the CSIRO to produce magnesium from magnesite mined at AMCÆs Kunwarara deposit. Over $200 million has been spent developing the AM process, designing, building and operating a demonstration plant at Gladstone and to complete an environmental impact statement and receive environmental approvals as well as funding for the commercial Stanwell Project. The paper outlines the development of the AM process technology, the operations of the Gladstone demonstration plant confirming the process technology, its operating costs and environmental performance. The paper covers the particular attention that was given to minimising the production of CHCÆs, the development of new cover gases for molten magnesium, the minimising of emissions of greenhouse gases and the life cycle analysis of magnesium as a replacement for steel and heavier metals in motor vehicles which confirmed its use resulted in a net reduction in global greenhouse gas emissions. The project received both Commonwealth and Queensland Government environmental approvals by August 2000.

Jan 1, 2002

By R Huntley, D Sims

Pajingo gold mine produces over 240 000 oz Au per year from underground mining using longitudinal bench stoping methods with a 15 m sublevel interval. Average mine head grade is around 12 g/t Au, contained within quartz veins between 0.5 to 16 m wide. The veins dip steeply (>60¦) and exhibit strike variations which lead to many curved and meandering drive/stope geometries. Unplanned dilution at Pajingo is predominantly controlled by fault planes that subparallel the vein and commonly lead to slabbing failure from the stope walls. Initial efforts at determining maximum spans were based on using the Stability Graph method of stope stability. However the method had a number of shortcomings that did not address mine site observations, including previous mining, interaction of different joint sets and the presence of discrete shears which resulted in poor prediction of stability. A method was developed to rate the risk of dilution for a stope based on observations made in the ore drives. The method utilises profile mapping of the ore drive, structural features observed and the orebody itself, on 10 m spaced sections along the ore drive. A matrix of factors determines the dilution risk for stoping at the time of mapping and the rating is transferred to longsection plots in the office. The dilution risk rating is used to determine the position and density of cable bolt support aimed at lessening the risk, and can be used to review blasthole designs prior to finalisation. The method has been used for more than two years and has proven to be a good indicator of potential unplanned dilution (ie waste outside of ring design). A recent review of the dilution risk mapping process has indicated it is assisting in reducing the impact of dilution at the operation and has the scope to be utilised further in the design process.

Jan 1, 2002

The issue of flyrock is critical to the operation of all mineral extraction sites that use blasting. This paper takes a risk-based approach to identifying, analysing and managing the flyrock hazard. The basic cause of flyrock generation is a mismatch between the energy available and the work to be done. It should be noted that the energy available depends on the charge confinement. This mismatch can be caused by an abrupt decrease in rock resistance (geological weaknesses), overcharging, inadequate delays between holes and rows, inappropriate blast design and inaccurate drilling. The mismatch can come about from two sides; either too much energy (charge) for a fixed burden (work to be done), or insufficient burden (work to be done) for a fixed charge. The main mechanisms of flyrock generation are rifling, cratering, face bursting and secondary blasting. Each source mechanism has different characteristics in terms of vulnerable locations and each requires different control measures. However, all controls require a high level of quality control and assurance. A range of different approaches for managing flyrock risk and determining blast clearance zone dimensions are presented. A prescriptive consequence-based approach is discussed. A conventional risk matrix based approach is outlined with some innovation regarding controls rating and uncertainty ratings. A published semi-empirical approach is discussed. Two examples of quantitative flyrock risk analysis are presented; one involves bench blasting based on a published paper and the second is a cratering example developed by the author based on consulting projects. Finally, a stochastic approach based on published work is presented which simulates cratering superimposed on face bursting mechanisms. Also, a stochastic analysis undertaken by the author of the cratering example is presented. Effective blast emission management is required for two reasons; firstly to reduce and/or eliminate safety risks and secondly to manage public perception of blasting risks such as flyrock. The conclusion of the current research is that the wild flyrock risk can be estimated using existing techniques. Furthermore, both qualitative and quantitative risk management methods as discussed in this paper can be used for flyrock management for a given blasting situation. A combination of both is considered to be best practice.

Jan 1, 2007

By Yeliseyev AK, Shestakov AM

Comminution is the most power-intensive process in the technology for ferrous quartzite concentration. That is why improving comminution efficiency leads to the most appreciable economic effect on the concentration process. The trend towards reducing the lump size of crushed ore before charging the mill is considered to be of major importance in improving comminution efficiency. A considerable number of investigations have been performed in this direction. At the same time, a strong interest is also maintained in the increase of comminution efficiency due to optimisation of process taking place in the mill. Recently, investigations of the power intensity of the breaking process inside the mill have been carried out at the Krivoy Rog Mining Institute. It was ascertained that the routine conditions of comminution at a low mill ball charge ratio resulted in an unnecessarily high power consumption due to the force of impact of falling and milling bodies far exceeding the power required for comminution. For example, the power value for falling balls of 60 - 125 mm in diameter reaches 0.3 kNm, and the force and stress due to ball contact with the lining reaches 60 kN and 250 MPa respectively. These values exceed the force required for comminution by a factor of ten. Analysis of the mechanism of ore particle size reduction during comminution in the ball mill indicates that the size reduction is due to impact load, attrition load, and static load. The efficiency of size reduction is very low. During comminution in the ball mill, impact loads prevail and amount to 85 - 90 per cent of the total load. In this situation, not more than 10 - 15 per cent of destruction is due to attrition.

Jan 1, 1993

By Heyes GW, Orrock GI

The fine grained complex orebody at Elura contains galena, silver and sphalerite in a matrix which is mostly pyrite; sulphides account for about 90% of the orebody. A differential flotation circuit is operated to con- centrate lead and silver in the first section and zinc in the second. The metallurgical performance to date indicates the need for improvement. The lead section has been sampled in detail to comprehensively assess its overall performance and to define problems associated with lead and silver flotation. The substantial amount of pyrite in the feed dominated the perfor- mance of each stage of the separation. It was responsible for extensive dilution of all concentrates and, as a consequence, for very large circulating loads. The rougher concentrate contained 68% pyrite some of which was rejected through the three stages of cleaning to 39% in the final concen- trate. However, this was still sufficient to hold the assay of this product below 45% Pb. Size by size behaviour shows that flotation rates of the valuable minerals, as characterised by their rate coefficients, were low through- out the circuit. There was a significant deterioration in these rates between rougher and cleaner I stages which resulted in large circulating loads between these stages. There was practically no recovery of coarse valu- able minerals to the final concentrate. The dilution of the final concen- trate by non-sulphide gangue was not a major problem. If the recovery of the galena and silver in the +38¦m size range could be raised to that of the - 381Lm size range, galena recovery would rise from 78% to 86% and silver recovery would improve from 46% to 59%. The Pb grade of the final concentrate would remain the same.

Jan 1, 1988

This paper is directed at the problems flowing from Australia's capital gains tax (CGT) on the disposal of exploration and mining tenements both in Australia and overseas. It will seek to identify the issues and give some idea of the dollar consequences of decisions lacking forethought. The paper should not only be seen in the context of Australian tenements. CGT applies to the disposal of any assets by an Australian resident anywhere in the world. Offshore incorporated companies are not necessarily free from CGT risk where it can be shown that control management and control of the relevant company is in Australia. Also the disposal by non residents of Australian tenements are potentially caught by CGT. Some will ask what has taxation got to do with a valuer who merely determines the value of assets. These days no determination can be blind to the consequences. Whether the consequences can encourage a movement in value is another thing dependent upon the circumstances; but? A role of an experienced taxation consultant is to point out parameters within which associated professionals can work. The parameters will vary from case to case, but these when clearly indicated may be of help to persons such as valuers. When one comes down to it real estate salesmen are never wrong. They have a feel for the market and almost through self fulfilling prophesies achieve sales at their guestimates. Whether they could have achieved higher considerations become virtually academic in the process. Of course I am not suggesting that tenement valuers can be compared to real estate salesmen, or am I? What is the value of a tenement? Is it the DCF/NPV type of generated value based on a large number of variable factors (ie what one may call the fundamental value)? Or is it what another may pay to purchase that tenement if it were on the market at an instant point in time? How big is the market, the quality of the bidders, the size of their purses, the needs of the buyer for comfortable fit, the premiums for foreigners, the standard of the marketing, the industry intelligence level of the bidders, the obligation to farm/sell down? The list goes on endlessly. One question must be.asked-"Is a range of values acceptable for tax purposes". ie Would any tax consultant ever provide a valuation to the Australian Taxation Office that indicates three or more possible values ie

Jan 1, 1987

The CSIRO acid alumina employs the essential steps of digestion, hydrolysis and calcination to produce pure alumina from ores too high in silica or too low in grade for the Bayer process. The two latter stages have been described in some detail previously but the digestion stage requires separate dis-cussion since ores vary widely in composition and reactivity. Kaolinitic clays require. roasting to 8000C to render the alumina soluble in the 11 per cent sulphuric acid used in the process while some ferruginous laterites require roasting to reduce the reactivity of iron minerals. The iron content of pregnant solutions is kept at a low figure by digesting at temperatures where the solubilities of ferric oxide or basic ferric sulphates are small. A.two-stage, counter-current digestion is necessary to extract a high percentage of alumina from the ores and to pro-duce a basic pregnant liquor, with an S03:Al203 ratio of 1. 9 suitable for high temperature hydrolysis. The second stage of digestion, called the modification stage, treats slightly acid liquors with exces's of ore in the temperature range 1200- 1400C. The first stage operates at 1700-190 C to extract further alumina from modification residues, using recycled acid obtained from later steps of the process. Both temperature ranges are critical for economic operation and for proper con-trol of the iron content of solutions. In batch experiments, thirty minutes is regarded as a suitable retention time for digestion and modification; ores which require much longer than this are less suitable for the process. Impurities. such assilica, titanium and vanadium oxides are discarded in the residue along with iron oxide and do not contaminate the final product. Bivalent metallic sulphates such as MgSO4 can build up in the circuit to a very considerable degree without affecting. product purity.Comparison with other processes shows that the C. S. I. R. O. process is superior for control of contamination by iron, which is the major problem in acid processes. The pure intermediate product (3 A1203. 4 S03. 9 H2O) is also more convenient for cal-cination to alumina than the hydrated aluminium sulphate usually recovered in other processes. Detailed cost comparisons must await pilot-plant evaluation of the C. S. I. R. O. process.

Jan 1, 1963

The working of alluvial deposits has from all times been a popular form of mining, not only from the simplieity of its operations, but also, in its more primitive stages, from the possibility of carrying it on without large capital, without skilled lahour, expensive plant, long terms of preparatory work, or, in fact, any other requisite than a shovel, a dish, and a stout pair of arms. It is possibly one of the oldest industries, and probably, also, it has been one of the most profitable.In tbe Australian States the happy days of rid alluvial deposits, when gold was to be hltd for the shovelling, are now almost over. The lust for gold has caused almost every corner of the world to be ransacked, and the more easily earned wealth to he exhausted: for the most part there remain now those deposits only which are too poor to be profitably worked by means of the implements and methods of old times. But, though the deposits have become poorer, the keen inventiveness of man has devised means whereby such large quantIties can be treated at a cost so low that he can still make a profit out of them. In this article, out of the many means devised, three only will be referred to-viz., hydraulic sluicing, bucket dredging, and hydraulic dredging. All mining may be divided into three principal operations-1st, the breaking down and conveying of the ground to where it is to be treated for the recovery of the metallic values, 2nd, the treatment of the material for the recovery of those values; and, 3rd) the removal of the worthless material or tailings. In all mining the object is to effect these three processes for the least expenditure of money commensurate with the highest degree of efficiency.

Jan 1, 1907

By Johnson G. D

The typical management systems used to address the issue of operating risk centre around technologies that identify, analyse and evaluate the risk factors. These logical systems such as HAZOPs, Fault Tree Analyses, etc rely heavily on technical considerations and process flow analysis. However, several research projects, across industries and countries have clearly shown that up to 80 per cent `operating risk' is determined by the actions of workers, supervisors and managers. Yet, most of the focus of risk management is on technical and procedural issues. This suggests that a significant imbalance exists in our management systems and that industries are unlikely to show significant improvement on current levels of safety performance. Effective risk management must include the intangible element of the organisation psychology towards risk. A practical and theoretically sound approach had been researched and developed in and for the mining industry. This paper describes the approach and methodologies applied at the Cosmo Howley Gold Mine in the Northern Territory, where an inclusive approach to managing financial and operating risk was followed. The Cosmo Howley Mine showed a dramatic reduction in lost time injuries and improvements in financial indices - providing conclusive evidence of the link between management effectiveness, safety and productivity. This `triangle' has often been the subject of speculation and theory. It is demonstrated in this paper how the management team of Cosmo Howley focused their practices and systems on the improvement of organisational dynamics, ensured tangible and clear focus on parameters of costs and safety and how the improvement process was managed. It is concluded that the improvement process did not require a `major re-engineering' of the organisation. The organisation culture changes that occurred were incremental, planned and effective. It entailed the implementation of sound risk management methodologies, and ensured that a significant emphasis was placed on an aspect that is defined as `organisational dynamics'.

Jan 1, 1994

A static equilibrium mathematical model of the lead,flotation circuit at The Zinc Corporation Limited is developed to permit the feasibility of applying computer control techniques to be studied. The model is based on the relationship between the cumulative recovery of a mineral species R from n flotation cells or assemblies in a unit given asR = M(1-(1 - p)?)where M is the maximum possible: recovery and p is the probability of flotation for the particular mineral.The verification of this relationship by data from pilot plant cell-to-cell and plant open flow flotation units shows that these systems behave as if the probability p is constant for each cell or assembly of anyone unit. It is also concluded from this study that mineral recovery is independent of the head assay. Two consequences of this probability relationship, are demonstrated.1. The total recovery from an open flow unit is enhanced by increasing the number of flotation assemblies but maintaining a constant flotation time and total air addition.2. The rougher and scavenger capacity required to recover a bulk lead-zinc concentrate is approximately half that used for selective flotation.The static equilibrium model is derived by solving a series, of mass balance equations which incorporate the recovery-probability relationship for each type of flotation unit. It is used to calculate the complete metallurgy of the circuit for any arrangement of cells knowing only the plant feed assays and the probabilities of flotation for the three mineral types, galena, marmatite, and gangue for rougher, scavenger, and cleaner flotation. The high degree of prediction obtained is demonstrated by comparison with plant results.Pulp flowrate is the principal variable affecting the accuracy of the present model. Work is in progress to establish this effect and permit the derivation of a dynamic model. Even so, the static model is valuable in its present form for the analysis and design of circuits, and optimisation studies using computer services have been conducted.

Jan 1, 1964

The Electricity Commission of New South Wales is the largest single user of coal in the State, burning some 8.5 million tonnes during 1972-73. This coal is converted in power stations into electrical energy which is then distributed throughout the State and supplies the electricity requirements of 99% of users in New South Wales. Barring any major change in energy consumption growth, or change in the current mode of power generation, the consumption of coal by the Commission is estimated to have a doubling period of from 10 to 14 years. Coal is used to generate up to 87% of the electrical energy sold, the balance being obtained from hydro gen- eration, mainly in the Snowy Mountains, and, to a much smaller extent, from diesel generation at Broken Hill. A very large proportion of this coal is consumed in pulverised fuel boilers in the Commission's power stations, located mainly on the various coalfields of New South Wales. Coal supplies are drawn from a large number of mines, some owned or controlled by the Commission, others being owned by public companies. The coals used, originate from at least ten different seams located in the northern, western and southern coalfields of New South Wales, and therefore have a wide variety of properties and characteristics. To ensure an adequate fuel supply well into the future, the Commission is continuously engaged in coal prospecting in various areas. Bore data is used in assessing coal reserves and is also the basis of economic studies for their possible future exploitation. The numerous coal cores obtained during drilling operations are carefully tested and analysed in order to assess their suitability for power generation purposes, and to provide data for the design and operation of boiler plant using these coals. In existing power stations coal deliveries are systematically sampled and tested for quality control, for pay- ment of purchases based upon certain quality parameters and to check for any possible plant operating or main- tenance problems that might arise from the use of the particular coals. Almost all the properties of coal that may be necessary in the Commission's work are obtained using standard testing procedures carried out by the Commission's chemists, in laboratories located at the various power stations or at its Central Laboratories in Pyrmont. Where specialised analysis or testing is required, this is ordinarily undertaken on a commercial basis by outside bodies, expert in the particular field. The significance of the various properties of the coal for use in p.f. boilers is outlined and related to their proper application to boiler plant design and to plant operation.

Jan 1, 1975

By D Morris

This presentation discusses a water filtration technology termed Fabric Media that is somewhat novel to the Australian market place and reports several recent mining industry installations. Like many novel household cleaning products, the Fabric Media technology applies the science of active washable micro fibres as the means of capturing contaminant solid particles. The simple Fabric Media water filtration product offers several attractive attributes when compared to the more traditional media bed filtration processes such as sand filters, and barrier filters such as micro membranes. A micro fibre based filtration media enables both superior adsorption for the contaminant and wash ability. Stronger adsorption capture enables a much higher flux and greater solids loading density for the filtration task, meaning that lower cost and smaller footprint equipment is available. For Fabric Media water filters, a textile weave is selected that minimises the flow backpressure for a given particle size separation. Thus for an equivalent filtration performance this provides a much lower operating energy requirement, generally between 50 and 80 per cent less than for a sand filter. A Fabric Media is not abrasive to itself or to the filter tank. The filtration performance remains constant and predictable. The equipment life is much longer and the asset value of the new filtration equipment should maintain throughout the project life. Fibres are extremely mechanically robust and chemically inert. Fabric Media survives repetitive exposure to a rapid mechanical washing event providing high wash efficiency with minimum equipment off-line time for the backwash. Dislodging captured contaminants by a high-speed mechanical agitation enables a minimum off-line time during backwash events. The vigorous backwash destroys adhering flocculant and eliminates the formation of mud balls, which can permanently foul sand filters. The mechanical agitation wash of the Fabric Media water filter ensures complete removal of grease and biofilm, which can adhere and block a membrane surface. In comparison to sand filters and micro filtration membranes, the Fabric Media product is applicable to a significantly wider range of solids concentrations without permanent damage. The technology is very forgiving to excursion in plant control and is therefore considered a very operator-friendly concept.

Jan 1, 2006

By W L. Lane, T R. Yanske

The Doe Run Company and its predecessors have been mining for over 130 years utilising the room and pillar mining method. This method has left approximately 20 per cent of the orebody in the remaining pillars. The original design and planning for the mines did not include extensive pillar extraction. This paper summarises the introduction of the pillar extraction and backfill techniques which have been implemented since 1991. These techniques include the use of low cost fill walls up to 20 m in height and the use of slinger trucks for topping off the fill. In conjunction with these accomplishments, much experience and confidence has been gained in the use of numerical modelling as a method of predicting pillar stability as the mining of pillars progresses. These advances have led to pillar extraction becoming a major portion of the overall production and have added to the mine life of Doe RunÆs six operating mines.

Jan 1, 1998

By D MacKenzie

The Rise & Shine Shear Zone is a shallow-dipping Cretaceous late metamorphic shear zone formed in a compressional environment. The Shotover mineralised system is dominated by normal faults with well-defined vein systems that cut across Miocene schist folds and faults related to the inception of the Alpine Fault. Despite the geological differences between these mineralisation systems, they share many common features of hydrothermal alteration that have resulted in disseminated gold in altered sheared schist. The most distinctive feature of both systems is the abundant ankeritic carbonate that has formed from decomposition of metamorphic chlorite. Excess iron from this alteration reaction combined with hydrothermal sulphur to form sulphides that host the disseminated gold. The chlorite decomposition reaction was reversed in the latter stages of both mineralisation systems, and this was accompanied by further gold emplacement in the Rise & Shine system.   Both alteration systems involved initial decomposition of metamorphic albite, followed by deposition of albite as a vein mineral in the latter stages. The fluid flow that yielded the disseminated gold was controlled by microshears that pervade host schists, in strong contrast to many Otago Schist vein systems where host rock alteration is absent. The ankeritic alteration in the Shotover and Rise & Shine mineralised systems has increased the volume of mineralised rock, and may have made suitable targets for large-tonnage, low grade gold deposits.

Jan 1, 2007