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Diesel-engine exhaust emissions represent a concern to the mining industry due to the poten- tially harmful health effects of the pollutants, as well as high costs associated with providing proper ventilation to assure a healthful working environment. The control of exhaust emissions is complex since the operating mode and condi- tion of the engine, the mine environment, and the operator's skills all influence the concen- tration and composition of the emissions. It is not practical to measure each exhaust pollutant in underground mines; therefore, a monitoring methodology is required to measure the pollut- ants and assess the air quality.

Jan 1, 1988

The development of the SECV brown coal open cut at Loy Yang will require depressurisation of aquifers both within and below the coal seam sequence in order to maintain open cut stability. Prediction of future open cut dewatering requirements at Loy Yang Open Cut and evaluation of the regional consequences of dewatering has required development of a detailed hydrogeological model of the Loy Yang area. The complex stratigraphy of the coal seam and aquifer/aquitard sequences at Loy Yang have been analysed and integrated with the hydrogeological concepts developed from regional groundwater studies to formulate a realistic hydrogeological model as a basis for finite-element flow modelling.

Jan 1, 1988

The Koongarra uranium deposit was discovered in August 1970 by Noranda Australia Limited. Denison Mines Limited acquired the rights of ownership in October 1980 and has continued the 'struggle' to bring the proposed mine, containing approximately 13 300 tonnes of U308 of Indicated ore and having a value of $1 billion (1984 dollars), into operation. During the 14 years that have passed, the owners have found it necessary to deal with numerous Federal, State and Territory government departments, aboriginal representatives, handle thousands of communications, attend and convene countless meetings and spend millions of dollars. This paper addresses the history, the requirements, the problems and the costs to date for what - no return.

Jan 1, 1984

Over the past 18 months the De Beers Consolidated Mines Ltd operations have made a concerted effort to move away from using the traditional shock tube initiating systems. These systems are being systematically replaced by the use of electronic delay detonators (EDD). Various trials were conducted in both host rock and kimberlite rock masses to improve tunnel advance as well as to optimise delay timing during trough openings (Ngoma et al, 1999; Mohle, 1999; Baltus, 2000). The high cost of EDDÆs, when compared with traditional initiation systems, led to a number of detailed studies being conducted on the mines where EDDÆs were being used. These studies aimed to quantify the additional benefits when blasting with electronic detonators. The studies showed that the change was justified on the basis of increased quality control and reliability gained through the use of EDDÆs. However, these benefits attract other related benefits, like fragmentation control, and backbreak reductions. When compared to the shock tube initiating systems the increased development face advance and the reduction of oversize during production blasting using EDDs, compared favourably to the less costly systems. As blasting engineers gained experience and confidence in the use of the system bigger blasts were initiated in mass under-cut blasts and slot raise development, by using multiple hole firing and second delay periods between holes. In the open pit and sublevel open stope mining methods the control of the fragmentation distribution and the effect of the mass of explosive detonated during a blast is detrimental to the loading and hauling production rates and the stability of the rock mass behind the blast. With a stable rock mass the bench cutting can be executed to establish steeper overall slope angles leading to large cost savings due to a reduction in waste stripping. It is the purpose of this paper to indicate through quantification that the use of the EDDs as an initiating system improves all-round blasting performances and assists in meeting customer requirements. The customer being the ore treatment plant.

Jan 1, 2001

By Chapman RL

A new process has been developed to remove the impurities copper, cadmium, cobalt and nickel from a zinc plant electrolyte by cementation on zinc dust. The process has been tested in the laboratory and in a continuously operated pilot plant. It has been found that cadmium, in the presence of small amounts of antimony in solution, assists the removal of cobalt. Cementation is conducted in two stages: the first stage for removal of copper with a small excess of zinc dust to produce a relatively high grade product, the second stage for removal of cadmium, cobalt and nickel by addition of zinc dust in the presence of anti- mony. The excess zinc and the cadmium present in the second stage cementate can be selective- ly recovered by slow leaching in spent electrolyte. Zinc dust processes in use elsewhere are described and compared with the present system for use in the Risdon circuit.

Jan 1, 1984

The Cleveland Mine has an annual ore production of 400,000 tonnes at a grade of 0.79 per cent tin and 0.30 per cent copper. It is achieved by an open stoping method, mechanised overhead sub-level benching. The operation is trackless with a ramp access from all working places to the surf- ace. Broken material is loaded from the stope and development headings into diesel trucks and hauled to the surface crushing plant.

Jan 1, 1977

The value of water is defined by its marginal utility. Challenges arise where the marginal utility is difficult to estimate. Therefore, when two or more interested parties wish to access water and they cannot use commensurate units for marginal utility; other means of value comparison must be sought. It is asserted here that sustainability principles can be used to guide this valuation if this is supported by a transparent decision framework. This paper proposes that the ICMM sustainable development principles provide appropriate guidance. In addition, a generic framework for decision making is suggested as a complement. Indicators of water management are recognised widely as an integral part of the accountability of minerals companies towards meeting sustainability goals. Less widely recognised are the ways in which sustainability principles can be used to guide decisions on water management which demonstrate growing awareness of its value. This paper illustrates the use of this logic in water issues associated with global sustainability challenges. It then goes on to demonstrate, using examples, that these links exist at all levels of the minerals industry and that indicators of this type are potentially as powerful as those currently being used. The reason for this is that the path from cause to effect is transparent and direct.

Jan 1, 2006

The paper describes the preparation of new rail guides, removing existing rails, and replacing with new rails in a vertical winding shaft.INTRODUCTIONThe general practice in coal winding shafts is the use of two cages, with one or more decks in which men and materials are raised or lowered. Winding speeds vary according to depth, the type of winding gear, and whether winding is done in downcast or upcast shafts. Each cage occupies a position between the centre and side of the shaft, with a clearance depending on the type of guides used. The function of the guides is to keep the cage in position throughout the entire depth of the shaft. There are three types of guides, two of which are in general use.1. Selected wood of about 6 in. square section, fixed to the side of buntons.2. Solid strand steel made up similarly to that of an ordinary wire rope, suspended from the head frame with weights attached to the lower end.3. Standard T-rail sections fixed to end or side buntons. Of the three types, 2 and 3 are in general use. The type used in the particular case, the subject of this paper, is T-rails of the standard section used in railway work.GENERAL DESCRIPTION OF SHAFTDetails of the shaft are as follow: Diameter of shaft: 22 ft. inside lining.Depth of shaft: 900 ft., including sump.Concrete lined throughout.VENTILATIONThe ventilation shaft is used for the upcast, through which all vitiated air passes to the main ventilation fan. Quantity of air passing up the shaft is 400,000 cu. ft. per minute. The air is highly saturated during the summer weather and slightly less...

Jan 1, 1947

The Goldfields extend from Norseman in the Eastern Goldfields to Marble Bar in the Pilbara; a distance of approximately 1,500 km. (Figure 1). Tailings dams are conspicuous on the relatively flat and featureless Goldfields' landscape. They impound wastes from mining and processing operations. Tailings dams interact dynamically with their environment and progressively consolidate as water is removed. Water recovery systems have been used over the last decade to improve water conservation, increase dam stability and to extend the service life of the dam. Without water recovery, dams would need to be significantly larger to permit the tailings to dry back sufficiently to support subsequent depositions. The paper reviews inter-relationships between the tailings dam and its environment and discusses water recovery systems, that are commonly used. The effectiveness of water management is illustrated by reference to seasonal water recoveries measured by Western Mining Corporation (W.M.C.) at their number 2 dam at Kambalda.

Jan 1, 1988

By Sun Yubo

A basic method for designing reasonable jig cycles for movable-screen jigs is presented in this paper. The reliability and effectiveness of the designed jig cycle was verified with tests performed on an artificial mixed material. An iron ore (12 - 100 mm) from Huashugo Iron Mine, Jiuquan Iron and Steel Company, was treated in the TJD-75 movable-screen jig using the designed jig cycle. A good result was achieved. The procedure presented here should provide a reliable technique for the design of a commercial movable-screen jig.

Jan 1, 1993

By D V. Bright

Oroya Mining LimitedÆs Mt Gibson Gold Project, 280 km north east of Perth, is one of Western AustraliaÆs more recent discoveries of a major gold resource. Modern gold mining commenced in 1986, on a shallow laterite reserve of 6.62 million tonnes averaging 1.50 g/t gold. By the end of 1998, mining had progressed from an initial three shallow laterite pits to a series of 14 open pits with maximum depth of 100 m, on three separate trends with a combined strike length of over 10 km. The Project has yielded 868 400 ounces of gold from 16.5 million tonnes of ore at an average recovered grade of 1.64 g/t gold, from near surface laterite, supergene and primary sulfide ores. This includes a 160 m deep underground decline operation which mined 120 000 tonnes at an average grade of 9.1 g/t gold and a four million tonne Dump Leach on a synthetic base which yielded 68 900 ounces. The Project was placed on Care and Maintenance in 1999 pending drilling to refresh reserves. Hosted by strongly tectonised and altered Archaean volcanic rocks within and adjacent to a major regional shear, primary gold mineralisation occurs as structurally emplaced sulfide and quartz lodes. On the main Gibson Trend, multiple zones of mineralisation occur within a zone up to 600 m in overall width. The overall footprint of known mineralisation exceeds 20 km2 and the potentially mineralised domain exceeds 100 km2. The current estimate of a combined resource of 9.4 million tonnes averaging 2.47 g/t gold (749 500 ounces) is derived from a database of 10 500 drill holes, of which only four per cent test below 100 m vertical depth. Ore envelopes which remain open at depth and major gaps in the drilling pattern beyond the currently defined resources promise straightforward potential to define additional resources. Oroya Mining Limited and its joint venture partner Barrick Australia are targeting a major gold mining operation from this potentially world-class mineralised system. Oroya has the exclusive right to produce up to 500 000 ounces and is working toward the completion of a feasibility study into a resumption of mining using the existing 1986-built 1 Mtpa processing plant. The joint venture is in the early stages of testing the depth and lateral potential of the Mt Gibson mineralised system. Mining options currently under review include progressive open pit extension and amalgamation leading to one or more large open pits, and underground operations. It is too early in the development of the Project to offer definitive economic models of potential large-scale mining operations.

Jan 1, 2003

Milling Practices Ore is direct tipped into a gyratory crusher and crushed to minus 209mm and fed to stockpiles for the blending of grade into the mills. Secondary and tertiary crushing reduces the mill feed to minus 12mm. Mill feed is a blend of breccia and porphyry ore into two primary mills with the overflow to a regrind mill and then into the leach tanks, followed by a conventional counter current carbon in leach circuit. Underflow from the cyclones feeds back into the primary mills for additional grind. An unusual feature of the Mt Leyshon milling circuit is a desire to produce scatts. The scatt by product is the polished base rock of the Mt Leyshon Breccia clast and contains little grade. These scatts are produced because of the additional time and energy required to produce a proportionally small quantity of gold. The residence of these scatts in the mills decreases the mill feed rate. Scatts account for twenty percent of the mill feed and are screened from the discharge of the primary mills.

Jan 1, 1990

The relationship between the permeability of coal to a mixture of methane and carbon dioxide, and the permeability to methane and carbon dioxide separately is described in this paper, along with the analysis of the differences in permeabilities of coal to methane and carbon dioxide.

Jan 1, 1989

By R Crookes

Mine to mill optimisation is not a recently developed concept. It is one which is finally gaining greater acceptance in the Australian mining industry, possibly as a consequence of the downturn in markets and certainly made possible by the development of tools and techniques to measure and predict performance of mining equipment and milling plants. Fragment size after blasting has a major bearing on the performance of all aspects of mining and a major influence on the economics of the mine and mill. The process to optimise fragmentation consists of identifying the key factors, measuring their sensitivity through a modelling and trialing program and conducting economic evaluations on the changes on both the mine and mill. Management of the Ernest Henry copper-gold mine in Northern Queensland, Australia, has espoused the mine-to-mill concept and is actively pursuing excellence in blasting to drive down the total cost of mining. Ernest Henry has extremely hard, wet and heavily jointed geological conditions, and had high powder factors and a large percentage of oversize material reporting to the crusher or ROM stockpile. A blasting improvement program was undertaken between Ernest Henry Mining, Thiess Contractors and Orica Explosives to optimise the ore size distribution, particularly the oversize component, as a step to reducing production costs. Measures of success included bringing the rock breaker back from double shifts to part-time operation and increasing direct dumping into the crusher from 70 per cent to 92 per cent. Tools and technologies to predict and measure the performance of individual blasts are needed before it is possible to optimise fragmentation or the downstream processes. These include such specialised Orica Explosives tools as SABREX, POWERSIEVE, and POWERLINE and commercially available drill monitors. As important are the more common still and video camera, borehole loggers, measuring tapes, surveying equipment such as face profilers, and a high speed camera if possible. There are many tools which can be applied however it is rarely necessary or possible to use all of them due to time constraints and a point of diminishing returns. The critical point is knowing what the true parameters of the blast were compared to design -only then can the predictive tools be effectively utilised. Stage one of the program has seen gradual improvement as the understanding grew and practices have been implemented resulting in reduced bridging in the crusher, geological mapping with respective blast designs and the oversize component under control. This paper is about a process approach which links tools and technology to understand the fragmentation drivers and then to optimising the size distribution. The continuous improvement program continues . . .

Jan 1, 1999

The characteristics'of the Darling Range bauxite deposits are compared with tho"se of other large deposits in Australia and over seas The Bayer Process is them described with particular reference to the way.in which the process has been adapted to suit Darling Range bauxite. The.main characteristics considered are the physical aspects of the deposit.and its' effect on mining techniques, the grade of the ore and its effect on residue treatment, the nature of the alumina, iron and silica. content and their effect on process methods, equip- ment and raw material. Usage.

Jan 1, 1973

Effective pre-splitting is a key feature of wall control blasting. Most mining operations use historical designs or rules-of-thumb for designing pre-split blasts. If productivity and safety benefits are to be gained, these methods need to be improved in order to accurately predict pre-split performance. This will become more pertinent as overburden depths increase, placing more demands on the pre-split operation. A classical qualitative quasi-static description of pre-splitting (Kutter and Fairhurst, 1968) is presented and it is shown to give some useful insights into the pre-splitting process. In particular, it is clear that the time interval between the initiation of adjacent pre-split holes is a very important factor. In terms of this analysis, it is shown that current techniques which employ detonating cord as the surface initiation system are not optimal. A static quantitative model for pre-split design which has been widely reported is then discussed and its simplifying assumptions and their implications are examined. It is shown that several of the assumptions may be reasonably justified in the light of real pre-split results. It is concluded that the model should yield conservative designs. For this model, the key calculation of the blasthole pressure that is needed to determine the required explosive charge is highlighted. Popular methods employed in the literature for describing the expansion of real detonation products are shown to be inappropriate and an alternative method using the non-ideal detonation code, CPeX, is proposed.

Jan 1, 1995

By Erceg MM

The Red Dome deposit is located about 15 kilometres by road west of Chillagoe in far north Queensland. The deposit is within a well known band of mineralisation which is adjacent to the Chillagoe - Mungana railway line. Earlier this century several mines were operating in this area, some of which were supplying feed-stock to the copper smelters which were operating in Chillagoe at that time.The area around Chillagoe has long been an interesting geological region with numerous locations exhibiting significant gold and base metal mineralisation.The Red Dome area of some time, however it was not became involved that the apparent. The upsurge in seventies added further investment.mineralisation had been known for until Noel Adam, a prospector,true potential of the area became the gold price in the late encouragement for additional

Jan 1, 1986

By Beattie T

For many years explosive sensitized explosives were assessed using the ballistic mortar. The standard test used lOg of explo- sive initiated with a detonator. Over the past 30 years many non- explosive sensitized explosives and blasting agents have been developed. Most of these new explosives and blasting agents cannot be tested in the ballistic mortar as they are not initiated by a detonator. It is also known that all explosives have a criti- cal diameter. When the cartridge diameter is less than the criti- cal diameter a stable detonation cannot be produced and the explosive fails to propagate. For many blasting agents the criti- cal diameters are greater than 50mm and such large diameters cannot be tested in the ballistic mortar. The underwater test allows explosives to be tested in diameters greater than the critical diameter and to be adequately initiated with suitable boosters. The effects of the explosive on the water are easy to measure and the test provides a safe, convenient and reproducible means of assessing explosive performance. The reaction products are allowed to expand to a low final pressure. The underwater test can therefore be used to give a measure of the total expansion work of explosives. The detonation of an explosive induces a shock wave in the surrounding water. The profile of the shock wave can be meas- ured 3-6m from the charge and the energy in the shock wave cal- culated. Measurements over a range of distances show that the shock wave is attenuated as it travels through the water and the attenuation is different for each explosive. This makes it difficult to estimate the energy dissipated in the water between the charge and a single gauge. The reaction products continue to expand and displace water. The bubble continues to expand until the kinetic energy of the water is zero. The gas pressure in the bubble is now lower than the hydrostatic pressure and the bubble radius is a maximum. The higher external hyrdostatic pressure causes the bubble to implode. The bubble energy is the work done displacing the maxi- mum bubble volume against the hydrostatic pressure and can befound from the bubble period. It is not always easy to find a large enough pond for the bubble to expand undisturbed by the water surface and the containing boundaries. This makes it difficult to determine the bubble energy accurately in shallow Ponds.

Jan 1, 1988

Bulk mining and handling technologies have been affected in two ways by the soaring costs of oil products in recent years. Firstly, the substition of oil by coal, oil sands and oil shale demands the development of high-capacity mining and handling equipment, and secondly the rising fuel costs favour a trend away from cost-intensive heavy truck transport. To cope with the growing volumes of material in mining and handling, larger equip- ment units have to be designed. This tendency becomes apparent from the examples of the Hambach open pit mine in Germany, and the Port Kembla handling plant in Australia. A combination of the flexibility of heavy truck transport and the cost advantages of belt conveyors is achieved by the system of semi-mobile crushing installations described in this paper.

Jan 1, 1981

By Ormsby W, Peters M, Van Gils H, Kehprens P, De Boorder H

The northern margin of the Arunta Block is characterized by a metasedimentary and volcanic sequence of lower metamorphic grade than that of the central Arunta Block. Whereas the Harts Ranges have granulite to amphibolite metamor- phic grades, the northern margin only reached greenschist to lower amphibolite facies. Most of the high grade area of the central Arunta Block appears to be unprospective, but the lower grade marginal fringe contains most mineral pros- pects to name a few: Barrow Creek, Molyhill and the Jervois Mine. Numerous smaller prospects occur in this marginal zone. Four major lithostratigraphic units can be recognized in the Jervois Mine area: 1. Gneissic Suite consisting of granitic and rhyolitic gneisses which are characterized by a distinct intersection lineation. 2. Metasedimentary Suite of cordierite- biotite schists and chlorite-biotite schists. 3. Mine Sequence, consisting of spotted schists with andalusite and cordierite, quartz- sericite-feldspar schists with distinct rhyo- litic remnants and quartz-sericite-feldspar schists which contain the lode horizons:

Jan 1, 1984