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The mining industry was one of the first to be directly regulated by legislation in the 19th century when British public opinion was mobilised by reformers seeking social change. In the next wave of mining safety legislation, account was taken of technological advances in the industry, but again, the amendments were in large measure driven by public expectations following a series of disastrous mining accidents with great loss of life. The result was the prescriptive and detailed type of regulation that has, to a great extent, remained with and become a way of life in the mining industry of today. In 1970, the then British Government appointed Lord Robens, a former Chairman of the National Coal Board, to head a Committee of Inquiry to consider the arrangements then in place for preserving the safety and health of people at work and to consider and make recommendations on the need for change in the legislation governing such matters. The result was the 1972 report Safety and Health at Work which, in turn, resulted in the Health and Safety at Work etc Act 1974 (UK). This Act, which embodied the now familiar concepts of consultation with the workforce on matters relating to health and safety and the statutory application of the previously existing Common Law concept of the duty of care, became the model for much occupational health and safety legislation in many places around the world, including the various States of Australia. As a result of the `Robens style' of legislation, the regulators of the mining industry have had the opportunity to become less prescriptive and to introduce an `enabling' style in their statutes. This has not always been welcomed by the industry which has, to some extent, missed the comfort of knowing exactly what is required of it in order to ensure compliance with the legislation while appreciating the `self-regulatory' opportunities that the new style of regulation offered.

Jan 1, 1995

Permian ultramafic and associated rocks form a belt in East Nelson and contain widespread, but minor, chromite and copper mineralisation and some mining was carried out last century. Since mining ceased, about 1910, little mapping of the belt has been carried out although its origin has been debated vigorously. Recently it has been inferred to form part of a ophiolite sequence, which has been tectonically disrupted to form a melange, and several 'plate -tectonic' models have been proposed. Mapping in the central part of the belt has shown that mTlanges are to the east of the ophiolite sequence. Faulting is widespread so that in anyone section part of the sequence may be absent. Some of the faulting is difficult to recognise and probably took place during or soon after the sequence accumulated. Other faulting is clearly of Cenozoic age and includes major NNE striking faults, with large horizontal displacement, and approximately east-west striking faults with almost entirely vertical movement. The writer agrees that 'plate tectonics' could provide an acceptable explanation for the origin of the ophiolite sequence but the models so far proposed do not take into account the nature and field relationships of the rock units present. Copper Mineralisation is largely controlled by the younger faults, particularly those striking about north-east, and is largely confined to the western part of the ultramafic rocks particularly where slivers of the volcanics of the adjacent Lee River Group are enclosed. Mineralisation, along NNE striking shear zones, has recently been found in the volcanics some distance from the ultramafic rocks. The chromite mineralization is in the form of primary layering in the basal part of the ultramafic belt although in places it has been concentrated into lenses probably during serpentinisation.

Jan 1, 1977

Acclaimed as one of the most remarkable orebodies anywhere on account of its length and richness, the Oroya Shoot occurs on the eastern side of the Kalgoorlie gold mining district close to the contact of a fine-grained with a coarse-grained greenstone, and in association with sediments. In the same stratigraphic position on the synclinal flanks of a large drag fold there are a number of other flat ore pipes, smaller, but almost as rich in tenor. Although differing in control structure they are believed to be related genetically as all show a characteristic development of a bright-green ore known locally as "green leader". Folding and bedding structures have played an important part in formation of the ore shoots.INTRODUCTIONThe Oroya Shoot is the biggest individual bonanza of Kalgoorlie's Golden Mile. It has been stoped almost continuously over a length of 4,300 ft., and about 2,000,000 oz. of gold have been won from it. Naturally pillars are not left in the best sections of such an orebody and the detailed geology of stoped out portions can only be inferred from examination of adjacent workings.No mention is made in the literature of the actual makeup of the orebody although a few structures are shown on some of the old cross sections. A diagrammatic cross section, Fig. 1, adapted from Stillwell (1929) illustrates the form of the Oroya Shoot.Gustafson (1937) drew attention to the role of folding, previously unrecognized, in Kalgoorlie geology, and indicated that the Oroya Shoot is associated with a minor drag fold on the east limb of the Brownhill Syncline, itself part of a larger left hand drag on the Central Anticline. The latter in turn appears to be part of a regional right hand fold system as shown by Campbell (1953).

Jan 1, 1959

Limnological processes are often overlooked when trying to assess the long-term water quality of pit lakes. It is essential to deal with these aspects, since they determine the geochemical and redox conditions around the submerged reactive rocks. Underground mining at East Sullivan (northwestern Quebec, Canada) stopped in 1966. Two adjacent stopes, 91 and 106 m deep, were left open to the surface, with roughly circular apertures of about 7700 and 18 000 m2. Water filling of the two glory holes went on over three decades and overflow began in 1999. Surface water pHs values were near neutrality as late as 1994 (6.3 - 7.0) but had become acidic by 1999 (2.6 - 3.5). Limnological processes were investigated in late-1999 and early-2000, with temperature measurements at 1 m intervals and sampling of the water columns before and after the late fall density inversions. A strong chemocline-thermocline was observed at 15 m. Above the chemocline, the water had low pH (3.2), no alkalinity, and low Na, Ca and sulfate. Below the chemocline, the concentrations of these parameters and the pH increased, with pH at 5.5 - 6. Consistent with the low pH values present in the upper 15 m, the concentrations of Al, Cd, Cu, and Zn were high near surface, and decreased below the chemocline. Dissolved oxygen concentrations as well as Eh decreased with increasing depth. Alkalinity integrated over the entire water column below the chemocline was at least six times larger than the acidity inferred from Fe2+ and Fe3+ concentrations. The temperature survey showed that seasonal density inversions do occur in the water column. The inversions do not involve the entire water body, since the surface to depth ratio is too small for a complete overturn. Oxygen from surface waters is carried to the bottom where it oxidises sulfides, until its depletion. Residual Fe2+, stable under anoxic conditions, is displaced upward by the next dense water incursion. It eventually reaches the thermocline, but only after many successive incursions. There, its oxidation to Fe3+ allows hydrolysis, which consumes dissolved and wall-rock alkalinity. It is only in 1999 that alkalinity was finally exhausted and that the pH shifts to acid values. The low pHs in the surface waters allow a build up of Fe3+ and Al3+, and desorption of metals from colloids. Buffering is assumed to be controlled by wall-rock oxyhydroxides in surface waters, and by carbonate minerals in the bottom waters. Because the seasonal overturns are far from being complete, the large amounts of alkalinity acquired from the wall-rock in the bottom waters cannot balance the acidity of near surface waters. Hence the surface water is acidic.

Jan 1, 2003

By D C. Curry, M F. Young

Many flotation operations need to improve liberation to increase recovery, yet often plant operators are concerned that finer grinding will produce æslimesÆ with low flotation recovery. Yet simple circuits can be designed to give good liberation and excellent fines recovery. The economics of fine grinding has changed fundamentally with new stirred milling processes like the IsaMill. This has also transformed flotation design due to the clean surfaces produced in the inert grinding environment and the narrow size distribution produced by the IsaMill. Past poor experiences with æovergrindingÆ in flotation circuits was often due to the wide size distribution and the degradation of mineral surfaces produced by conventional devices using steel media, like ball mills and tower mills. The correct practice is to design both grinding and flotation together to suit the liberation characteristics of the ore. This may result in several stages of grinding and flotation, recovering mineral as soon as it is liberated, then applying further grinding energy only to those particles that still need liberation. Flotation then occurs in relatively narrow size bands and with minerals with similar liberation characteristics. Though the circuit may look more complex on paper, in reality it is far simpler to operate since each section performs predicably. Circuit design also needs to address materials handling. Finer particles affect froth handling and dewatering, but can still be easily handled by standard products with careful attention to design of launders, pumps and pumpboxes, and de-aeration before thickening and filtering. Excellent examples of designing circuits to include ultrafine grinding include Mt Isa and McArthur River (MRM). In the last decade these operations have produced over ten million tonnes of concentrate with an average grain size below 10 ¦m and above 80 per cent recovery. At Mt Isa, zinc recovery was increased by ten per cent by applying fine grinding to existing poorly liberated streams. In fact, grinding finer in the IsaMill actually improved the recovery of fine particles by ten per cent, and maximum particle flotation recovery occurs for 10 ¦m particles. A more recent example is the flotation of Platinum Group Metals (PGMs). The use of conventional ball or tower milling to improve liberation is severely limited by the impact of commonly associated chromite on classification, and degradation of flotation performance by the steel grinding environment. The ability of open circuit IsaMills to deliver a narrowly sized ultrafine product with clean mineral surfaces is transforming the design of platinum flotation circuits. The science and practice of fine-grained flotation is now thoroughly proven. The availability of the 2.6 MW IsaMill has made ultra fine grinding a practical tool for flotation plant design. It is expected to become a standard choice for future applications of both fine-grained ores and for minerals with difficult flotation characteristics.

Jan 1, 2005

Exploration risk increases with time, due to a general decrease in prospectivity as the more accessible deposits are found, and be- cause of rising costs and demands due to an improving standard of living. Risk is separated into financial risk, which increases sharply at the development stage, and the risk of failure which is at its peak at the outset of exploration. Risk is also divided into internal and external factors; some of the latter cannot be offset, viz. the systematic overall risk for any enterprise. The recently emergent risk embodied in withdrawal of large areas from exploration access is discussed. An appeal is made for objective multiple land-use policies to be applied so that land remains available at reasonable cost and with appropriate environ- mental safeguard conditions. Risks specific to gold exploration are discussed. As gold now accounts for over 50% of all mineral exploration, and most other metals are depressed, any sharp decline in gold will have grave effects on the ex- ploration industry. Gold exploration will become more difficult and expensive in about five years, after most of the readily accessible deposits in Australia are found. Government's role is important in pro- viding stability and basic data at reasonable cost. Exploration for specific minerals or of particular areas could be stimulated by extending the 150%.tax deductibility of R & D to aspects of exploration of strategic im- portance, particularly in the event of a downturn. Exploration is classic scientific research, at least through to the proving Stages.

Jan 1, 1987

By M P. Schwarz

In recent years, computational fluid dynamic (CFD) modelling of mechanically stirred flotation cells has been used to study the complexity of the flow within the cells. In CFD modelling, the flotation cell is discretised into individual finite volumes where local values of flow properties are calculated. The flotation effect is studied as three subprocesses including collision, attachment and detachment. In the present work, these subprocesses are modelled in a laboratory flotation cell. The flotation kinetics involving a population balance for particles in a semi-batch process has been developed. From turbulent collision models, the local rates of bubble-particle encounters have been estimated from the local turbulent velocities. The probabilities of collision, adhesion and stabilisation have been calculated at each location in the flotation cell. The net rate of attachment, after accounting for detachments, has been used in the kinetic model involving transient CFD simulations with removal of bubble-particle aggregates to the froth layer. Comparison of the predicted fraction of particles remaining in the cell and the fraction of free particles to the total number of particles remaining in the cell indicates that the particle recovery rate to the pulp-froth interface is much slower than the net attachment rates. For the case studied, the results indicate that the bubbles are loaded with particles quite quickly, and that the bubble surface area flux is the limiting factor in the recovery rate at the froth interface. This explains why the relationship between flotation rate and bubble surface area flux is generally used as a criterion for designing flotation cells. The predicted flotation rate constants also indicate that fine and large particles do not float as well as intermediate sized particles of 120 - 240 ¦m range. This is consistent with the flotation recovery generally observed in flotation practice. The magnitude of the flotation rate constants obtained by CFD modelling indicates that transport rates of the bubble-particle aggregates to the froth layer contribute quite significantly to the overall flotation rate and this is likely to be the case especially in plant-scale equipment.

Jan 1, 2005

This paper summarises the performance of the gold exploration industry in the Western World since 1970, with particular emphasis on the period 1985 - 2003. It identifies the number and size of major (ie >1 million ounce) primary gold discoveries made over this period by location and type of company. Estimates are also made of the likely contribution from smaller discoveries down to 0.1 million ounces. It also assesses the overall trend in the percentage of grassroot discoveries that are converted into mines, and the time taken to get them into production. By comparing the discovery data against total exploration expenditures it is possible to determine the average unit cost of finding a major gold deposit. It is significant to note that the average unit cost per discovery and per ounce of resource has tripled over the last 30 years. An attempt is made to determine the overall economic value generated by the industry. This is based on a detailed analysis of the rates of return and the Net Present Values (NPVs) of 109 major gold projects. After including an adjustment for discoveries not modelled, the NPVs are compared against the total amount of money spent on exploration to determine the value-to-cost ratio for the industry. Based on historical spot gold prices and tax rules, it is found that the return on investment for gold exploration in the Western World over the period 1985 - 2003 was around five per cent real after-tax. The returns achieved in the US and Australia was significantly higher than that for Canada. The returns also appear to have declined in recent years. With regard to the relative performance of junior versus major companies, the paper estimates that the junior sector accounted for 30 - 40 per cent of the total exploration expenditure, and were was associated with 34 per cent of the total number of major deposits, 27 per cent of the contained ounces and 25 per cent of the economic value found.

Jan 1, 2004

By S Coughlan

"A personal account of my journey from a student from suburban Melbourne in the 1970s through 40 odd years in and around the mining industry including the important events that have moulded my thinking and, as result, have formed the guiding principles of the mining contracting business and related entities that make up the Byrnecut Group. On reflection the game of chance has played a big part in the journey, from the selection of initial study based on my parents financial capability and the one and only offer of early employment by Western Mining Corporation where the formative phase of my career took place in what was the vibrant can do environment that existed in Kambalda in the 1970s and early 1980s. Subsequently the unplanned meeting with Frank Lubbock drawing me back to Kalgoorlie from Fiji in 1987 and the resultant formation of a private mining contracting business, the excitement of the mid 1980s Kalgoorlie mining environment followed by the turmoil of the 1987 stock market crash where business survival was the game, resulting in a relationship with the German based Thyssen Group.Thereafter the development of a large privately owned international multifaceted mining services group and the principles behind that growth. These include: development of strategic personnel and retention of those personnel cashflow management and debt within a private company scenario recognising opportunity for synergetic growth but insisting on market forces driving those business elements, as opposed to internal divisions group-wide sustainable business systems, particularly in relation to leading safety and training processes innovation and renewal to provide the elusive advantage in a competitive market the role of entrepreneurial small businesses in development of technology.The above principles combine to make a margin in a cut throat contracting environment, competing on the world stage against other contractors and owner mining options.CITATION:Coughlan, S, 2014. A miner’s journey, in Proceedings 12th AusIMM Underground Operators’ Conference 2014, pp 3–10 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Mar 24, 2014

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

A BRIEF account of the water-power potentialities of Tasmania is given, together with possible applications of the power commercially. A short history of the Waddamana development is then given, and the main works are described.WATER POWER POTENTIALITIESTasmania is an island of approximately 24,500 sq. miles area (exclusive of small islands). An area of 12,000 sq. miles is at an elevation of over 1000 ft. above sea-level, and of 2300 sq. miles at over 3000 ft. above sea-level. Average annual rainfall varies from' over 140 in. on West Coast to 20 in. on East Coast, and for the whole island averages 48.5 in. The total water-power available, based on conservative figures for run off and efficiency, is 6,000,000 b.h.p. Only a small percentage of this total will, under present conditions, and under likely conditions in the immediate future, be capable of being developed commercially.An extensive hydro-metric and water-power survey is at present being carried out in order to collect information concerning the most promising water-power schemes. Already 400,000 b.h.p. has been located in different proposals, and there is still a large area of very likely country along the West Coast which has not been investigated yet. Tasmania thus has very vast power potentialities, which, combined with good ports and communications and wonderful mineral and timber resources, should augur well for the establishment of the great key industries, particularly chemical and metallurgical.

Jan 1, 1923

Agitation leaching data has generally not been analysed to extract the maximum amount of information that is available. Analysis of heap leaching systems has been developing over the last five years to include the underlying rate determining process. This has been found to be diffusion control for all copper leaching; irrespective of whether they are oxide or sulfide minerals. Extension of the diffusion model has enabled better scale-up protocols to be developed. Acid consumption in copper heap leaching operations has been found to be a first-order chemical controlled system. The analysis of the data is relatively straightforward. However the interpretation and scale-up is not as direct as for agitation leaching. The differences in the residence time distributions (RTDs) are required to properly account for the scale-up between test columns and field. The leaching of oxide copper in agitated systems has also been found to be diffusion controlled. However, there are significant differences in the control mechanisms. Initially the leaching control is the supply of acid to the mineral surface. This is limited by the diffusion through the surface boundary layer and is thus subject to the level of agitation. Higher agitation provides thinner boundary layers and faster diffusion rates. Similarly, higher grade material creates a thicker concentration boundary layer and relatively slower leaching. Acid consumption in agitated leach systems is also a first-order system. However, the interpretation of the RTD is easier from batch testing to continuous multiple CSTR series plant operations. Altering the method of acid delivery can significantly reduce acid consumption rates. Predilution of the acid minimises the local acid concentration near the addition point and can reduce the consumption by up to 50 per cent. Modelling of the combined leach rate(s) and acid consumption rates in agitated systems allows comparison of various leaching strategies and selection of the most appropriate. Similarly, the model allows the identification of economic limits to leach recovery; where the cost of acid consumption is greater than the value of the copper recovered. From this analysis selection of residence times for agitation leaching can be determined. The modelling tools are useful in identifying optimum conditions for the leaching process.

Jan 1, 2005

By A L. Hemphill

Conventional models, based on pulp kinetics, were regressed directly to steady-state data from a nine-stage platinum flotation plant, to provide a basis for optimisation studies. æFroth factorsÆ were used to characterise the froth in each stage and a ædepressant factorÆ was used for gangue minerals, to account for addition of depressant to the cleaners. These factors improved the fit, but differences between plant and model of up to 50 per cent were still present in a number of streams in the cleaning circuit. A new model was developed, which incorporated two rate constants for the recycle of particles from the froth. These parameters provided a workable approximation of the distribution of probabilities of particles being recycled in plant cells, depending upon where particles appeared at the surface. The feed streams were characterised by two pulp kinetic parameters and the above recycle parameters, for all minerals and all stages. The fit to the data was improved very significantly, when compared to the use of conventional pulp-based models with several pulp rate parameters. This relatively simple model did not require additional data from batch tests and it was suitable for plant simulations.Simulations of various strategies for plant operation were tested, allowing the simulator to vary the concentrate flows from all or some of the stages, in order to maximise platinum recovery. The simulator was imbedded in an automatic optimiser and constraints on chrome content or mass flow of final concentrate were used. The simulations of the existing circuits demonstrated that an improvement in recovery, (for a constant concentrate mass) of about 0.5 per cent was possible, by optimising all concentrate flows. A significant reduction in the flow of scavenger concentrate and an increase in the recycle flows in the cleaners were noted. The results of simulations of various changes to the circuit configuration are also discussed. An increase in recovery of 1.4 per cent (for constant concentrate mass) has been achieved in recent simulations, through circuit changes. However, it was necessary to keep the mass pull of concentrate, per unit area, within typical limits.

Jan 1, 2005

By J Vann, S Dunham, M Job

Two-dimensional (2D) estimation approaches have significant advantages over three-dimensional (3D) approaches when dealing with a narrow vein- or layer-like geometry. The main motivation for 2D estimation is that bias due to non-additivity of grade is avoided. The various aspects of 2D estimation are illustrated by use of a case study from the 1A Shoot Nickel deposit at Leinster, Western Australia. The width of the 1A mineralisation is variable, and so are the angles of drilling intersection. Consequently, the variable of interest (in this case nickel grade) is not suitable for direct kriging because it is defined on non-uniform support. However, grade can also be defined as the ratio of two variables that are suitable for direct kriging: thickness and accumulation. In the 2D approach, tonnage equates to thickness and metal equates to the product of grade and thickness, also called æaccumulationÆ. In the simplest approach we can independently krige accumulation and thickness and derive the grade by the ratio of the two resultant estimates. The thickness and accumulation variables must be in a consistent projection orientation, for example a horizontal projection in the case of steeply inclined mineralisation like 1A. Because the density is not constant across the 1A (and most other) deposits, the issue of density weighting is also discussed. The impact of the correlation between grade and length, or grade and density are important issues in implementation of 2D estimation. The circumstances under which these correlations need to be taken into account (by co-kriging) are reviewed. Finally some comments are made on the final use of the resultant 2D estimate for mine planning purposes.

Jan 1, 2003

Pasminco is actively addressing the optimal utilisation of the known orebody at the Broken Hill mine, in addition to actively exploring for economic mineralisation in the region. As part of the policy of better utilising the known mineralisation a major upgrade of geological systems at the mine was initiated. The software chosen for this change was Maptek's Vulcan, mainly because of the ease of use of this package. Rather than having numerous data sources in varying software packages the mine aimed to limit interfaces and packages. The decision to upgrade the mine's software was not taken lightly, especially in view of the inature stage of Broken Hill mining operations. However, the current mine life of Pasminco Broken Hill mine is I I years, which is longer than many new mines setting up in Australia today. Also, a more sophisticated system brings with it a better understanding of the economic mineralisation, possibly extending the mine life further. The process of moving the Broken Hill databases into Vulcan was a huge task which included, database transfer, digitising, wire framing, modelling and eventually ore resource calculation. To ensure the validity of the new ore resource accounting system, the existing system was run as a duplicate for the introductory year, providing a comparison and if necessary a safety net. The benefits of such a change, however, became apparent almost immediately. Area geologists realised the opportunity to revise geological interpretations in their areas of responsibility, enabling them to obtain a better understanding of orebody morphology. Three- diniensional analysis provides an excellent method by which review of ore targets both in the underground and mine lease exploration areas can he made. As each new find of mineralisation enters the system and hopefully extends the mine life, the opportunities for further mineral exploration in Broken Hill area by Pasminco are enhanced.

Jan 1, 1997

By M Kilburn, A Edenton, M Adam, A Dehghan-Manshadi, R Smyth, L Lu

The blast furnace is a counter current reactor. The counter current principle of the furnace makes it very efficient in transferring the heat and mass required to drive the reduction reactions inside the furnace. However, this does impose a stringent requirement on the permeability of burden material in the furnace. Lump ore is one of the key ferrous feeds used in the blast furnace ironmaking process, often accounting for ten to 20 per cent of the blast furnace ferrous burden. As the lump ore descends through the blast furnace, it is dried, heated and reduced by the ascending hot reducing gas. To ensure good furnace permeability, generation of fines from burden materials including lump ore due to mechanical, thermal and chemical processes needs to be minimised. Unlike sinter and pellets, lump ore is subject to thermal degradation. To assess this behaviour of lump, the ISO decrepitation test (ISO 8371, 2007) is often conducted, and the result is calculated as a decrepitation index (DI). However, the DI values obtained from the existing ISO decrepitation test vary considerably between test portions and between laboratories, which makes meaningful comparison very difficult. In order to identify the sources responsible for the unacceptable repeatability, the effect of various testing conditions, such as particle size, heating rate, sample preparation and loading method, on the mean DI value was studied. The mechanisms contributing to thermal degradation in the top of the blast furnace and during the standardised decrepitation test are also discussed.CITATION:Lu, L, Kilburn, M, Dehghan-Manshadi, A, Adam, M, Edenton, A and Smyth, R, 2013. Thermal degradation of lump ore in the top shaft of blast furnaces, in Proceedings Iron Ore 2013 , pp 421-426 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 12, 2013

The purpose of this paper is to present a detailed account of the development of the Blasthole Diamond Drilling method of stoping now used at the property of Big Bell Mines Limited, in Western Australia.The ore body is one of the large low grade gold deposits of this State, being a continuous lens 1,850 ft. long and ranging in width from 15 ft. at the ends to 120 ft. at the widest part, the average width being 59 ft. It was first worked in 1913 by a local syndicate, and up to 1924, when operations ceased, 64,448 tons of 3.4 dwt. ore were mined from a shallow open cut. Between 1928 and 1932, 15 deep diamond drill holes were put down to test the ore body at vertical depths ranging from 100 to 700 ft. On the information gained from these drill holes, Premier Gold Mining Co. Ltd. began underground explQration. A level was driven at a depth of 253 ft. and crosscuts put out to the full width of the, orebody at intervals of 100 ft. along the strike, all ore broken being put through a sampling plant. On completion of this sampling,' Big Bell Mines Limited was formed as the operating company, and a plant with a capacity of 1,000 long tons per day was erected and mining and milling operations were commenced in 1937. Ore was first mined from glory hole stopes, 65 ft. long, separated by 35 ft. pillars, by benching using conventional steel with changes up to 20 ft., all broken ore being hauled to the main shaft on the...

Jan 1, 1948

By W van der Westhuizen

Mining professionals recognise the need and application for reconciliation and value tracking. It is necessary however to examine our current reality, and evaluate the following. Do our reconciliations really balance correctly, and are we able to track value effectively throughout processes? If we are honest, in most operations the answer is a resounding no. This means that we are surrounded by fluctuation and deviations that we do not account for and do not understand. This is astounding, because given our level of scientific knowledge and the technological tools at our disposal we should have a good understanding of our environments and be able to manage them effectively. How do we manage effectively if we are surrounded by so much uncertainty? The purpose of this paper is to describe, at high level, the principles of reconciliation and value tracking, to highlight some areas where deviations are introduced into the mining process, and to propose a different approach to the apparent reconciliation and value tracking conundrum. The paper also poses several questions that need to be critically addressed at most operations. Instead of making the same mistakes each month in reconciliations, mining professionals need to acknowledge two facts. Firstly there will be uncertainty that we need to accept, since all our estimates are based on sampling, an orebody is not homogeneous and peopleÆs behaviour will influence the process, etc. Variability will always exist and lead to uncertainty. What is the current approach in the reconciliation process when a deviation occurs and how do we use reconciliation and value tracking to manage and improve our business? The second fact is that we need to stop using reconciliation and value tracking to only measure and report. We need to adopt a new approach to manage mining throughput, where reconciliation and value tracking are enablers to the success of the business. In most mining operations, significant effort goes into planning and measurement for the purposes of reconciliation and value tracking. The importance of accurate and effective reconciliation and value tracking should be to ensure the maximum utilisation of the resource, maximising throughput and enabling the business to deliver a product to its client, of the correct quality and in a timely manner.

Jan 1, 2002

By P M. Downes

Porphyry-related skarns are common around thePacific Rim. Examples include Cadia and Goonumbla (Australia) Ok Tedi (PNG) and Ertsberg and Grasberg (Indonesia). Although usually treated as separate to the adjacent porphyry copper-gold (Cu-Au) deposit, skarns near porphyry deposits are commonly inextricably part of the larger Cu-Au system(s). Using Cadia as the primary example, we demonstrate that large, zoned, calc-silicate skarns necessitate very large porphyry systems. The Big Cadia and Little Cadia Fe-Cu-Au skarns are associated with the Late Ordovician to Early Silurian high-K calc-alkaline to shoshonitic Cadia Intrusive Complex (CIC). Together, these skarns total ~40 Mt of Cu-Au ore but are much smaller than the adjacent Cadia EastûFar East, Cadia Hill, and Ridgeway porphyry-related Cu-Au deposits. Historically these skarns were large, economically significant and highly controversial deposits. Porphyry-related skarns typically have well-developed calc-silicate and ore zonation. At Cadia, classic skarn zonation has been mapped between Cadia Quarry (proximal) and Big Cadia (distal). Over this 800 minterval, mineralogy trends from: proximal garnet > pyroxene, grading to; fe-oxide-rich skarn; and distal haematite-epidote-rich skarn. At Little Cadia, analogous but symmetrical alteration zonation is present. Big Cadia formed from fluids that migrated laterally within calcareous units and were sourced from strongly altered quartz monzonite phases of the CIC at Cadia Quarry. Isotope studies and mass-balance calculations suggest that prograde and retrograde skarn alteration was caused by magmatic dominated fluids. They also indicate that the mapped alteration system at Cadia Quarry is insufficient to account for the total volume of Fe present at Big Cadia. Despite the typically oxidised nature of many porphyry systems, not all skarns in the porphyry environment are oxidised. Assimilation of carbonaceous sediment by hydrous, alkaline intrusions could lead to reduced Au skarns rather than porphyry deposits. An EXTENDED ABSTRACT is available for download. A full-length paper was not prepared for this presentation.

Jan 1, 2008

This paper concerns a pragmatic approach to a problem that is common to many operations in the mining industry: how, over the short-term, to grasp the extent and distribution of long-term financial risk from environmental causes and reduce these risks strategically, in order to maintain stability and flexibility in the operations. Surprisingly, many companies operate without an environmental master plan or tools that can help managers measure and reduce financial risks. Such companies may be subject to unknown unacceptable financial risks from environmental causes. While many companies have some kind of internal system of environmental standards, it is often the case that such systems have uneven or limited effect, and offer no opportunities for visualising, measuring or managing aggregate financial risks from environmental causes. `Aggregate risks' are defined here as the calculated total expected financial costs to a company, or a subdivision of a company, from environmental causes at nominated operations over a defined time period. Aggregate risks take into account the type and magnitude of an environmental problem set, the likelihood of occurrence, and the time frame during which the problems will develop or solutions are required. For the large corporation, the 'amount of' aggregate environmental risk is a statistic that can be managed, along with other financial parameters, at the corporate level to improve the stability and predictability of operations, and protect the interests of shareholders. This paper concerns development of a management framework and a tool for the large corporation that responds to the need for managing aggregate environmental financial risks. The logical framework proposed to overcome the need for strategic vision is the Environmental Master Plan, and this is described in the next section. The analytical tool proposed to measure aggregate risks and provide relevant data and insights to managers, is called an Environmental Risk Management Information System, and its mechanics are described in a subsequent section.

Jan 1, 1994