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By Kallioinen J

This paper discusses different possibilities to enhance coarse particle flotation using forth manipulation and improved flotation cell design. Experiments were carried out in laboratory, pilot and industrial scale ranging from 0.0015 to 100 m3/axle. During the work it could be shown that the conventional idea of one single flotation cell type for both fine, immediate and coarse flotation of coarse and fine particles need different impeller tip speeds and a different froth control objective for effective operation. For optimal operation for coarse feeds the tip speed must be reduced and the froth vertical velocity must be increased to compensate the respective drainage velocity and also the water film thickness must be taken care of. Turbulent eddies below the froth interface should be suppressed. The paper discusses the effects of different froth manipulation variables like froth vertical velocity and froth residence time to the recovery and selectivity. A function has been developed, between the residence time of the froth and the flotation cell's ability to produce an improved concentrate grade as a function of different operating and froth manipulation variables.

Jan 1, 1993

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

Reconnaissance geochemical exploration in the vicinity of low angle detachment faults which bound the core of the Paparoa metamorphic core complex yielded Au up to 14 ppm, with 35% of greywacke, granitoid and quartz vein samples in the northern range exceeding the 0.01 ppm detection limit. Au is associated with Ag, As, Sb and Cu anomalies consistent with epithermal-type mineralisation. The Ag and Cu anomalies also distinguish this mineralisation from older, mesothermal Au deposits widespread in the Ordovician 'slate belt' greywackes, which comprise much of the upper plate of the core complex. Baryte, molybdenite, fluorite and stratiform U mineralisation in the northern Paparoa Range may also be related to hydrothermal activity along the detachment fault zones.

Jan 1, 1995

By Mauk JL, Arehart GB, Mathews SJ

The banded nature of the quartz veins which form the Empire lode at the Golden Cross mine, is perhaps their most striking feature. Petrographic studies show that significant differences occur between the bands in terms of textures, quartz grain size, calcite, adularia and kaolinite content, and the form, concentration and mineralogy of sulfide phases. The 818Oqu i values of individual bands within the vein material range from 7.3-10.9%x. Calculations show that the 818Oqäa,u values can be adequately explained by temporal variations in the system temperature, boiling and/or variations in the systems fluid/rock ratios.

Jan 1, 1995

By Yu H, Wan B

40 percent of the coal reserves and 40 percent of the ouput of coal in China is of thick coal seams. The sub-level caving in Yangguan has evolved together with the hydraulic supports, to provide recovery rates today of 80 to 85 percent, waste ratios less than 10 percent and daily outputs exceeding 7,000 tonnes. Matching of rate of shearing bottom coal to caving of top coal is most important to achieve "simultaneous" operations. Spontaneous combustion in the goat, gas, including in the goal, and dust, are particular safety hazards which are combatted with various stringent measures.

Jan 1, 1991

World class ore deposits can be defined commercially as deposits which contain metal concentrations in the upper quartile of resources size and/or grade, world wide, from which economic extraction can be undertaken, over a production period exceeding ten years, at a cost below the weighted average cost of cumulative production of the contained principal commodity of mines world wide.

Jan 1, 1993

Hamersley Iron Pty. Limited mines and pro cesses iron ore from two orebodies at Tom Price and Paraburdoo, situated in the North West of Western Australia. Conventional open pit mining methods are used at both sites in that material is drilled and blasted in benches, loaded by shovels and hauled to the Primary Crushers or dumped on waste or low grade material stockpiles. The present drilling fleet, consisting of eight high capacity Bucyrus Erie Rotary drills has evolved from an initial fleet of a large number of small percussion rigs. There are significant physical differences between the ore types encountered at both mines which nec- essitate the drilling of smaller diameter blast holes at Paraburdoo than those at Tom Price. The drilling practice at each site is dis- cussed, with examples of performance achieve- ments under varying drilling conditions.

Jan 1, 1978

By P Licence, M Stevens, P Atkinson, J L. Mauk

The Rahu and Ascot epithermal Au-Ag prospects are two km north of the Karangahake deposit in the Hauraki Goldfield, and all occur in an approximately 6 km2 north-south elongated magnetic æquietÆ zone. The prospects are hosted by andesitic flows, lithic-crystal tuffs and local sedimentary rocks; local silica sinter occurs at Ascot. These host rocks are intensely altered by quartz, adularia, albite, chlorite, pyrite, illite, interstratified illite-smectite and calcite. Several minerals show zoned distributions, such as widespread adularia that overlies more localised and deeper albite. Illite is the only dioctahedral phyllosilicate in several drill holes, but elsewhere it forms a shallow carapace underlain by interstratified illite-smectite with 50 to 90 percent illite. Veins are common with quartz dominating near the surface and calcite becoming more abundant at depth. Local laumontite-calcite veins also occur. Pseudosecondary and secondary fluid inclusions in quartz and calcite at Rahu and Ascot have homogenisation temperatures (Th) that range from 162¦ to 268¦C and 128¦ to 243¦C, respectively. Most inclusions have trapped dilute fluids with salinities of less than 1.9 weight percent NaCl equivalent, but 20 percent of the inclusions have salinities of 2.0 to 10.5 weight percent NaCl equivalent. Calculated paleowater table elevations suggest that the water table at Rahu occurred at 440 m above sea level (asl). At Ascot, silica sinter formed from discharging geothermal fluids (=100¦C) at the paleosurface at ~135 m asl. However the sinter also contains comb quartz with pseudosecondary inclusions that have a Th average of 229¦C, corresponding to a paleowater table at ~430 m asl during inclusion entrapment. Taken together, these results suggest that the Ascot sinter was buried by approximately 300 m of cover material during a major volcanic or tectonic event, and that subsequent hydrothermal fluids formed veins and alteration at Rahu and deposited comb quartz at Ascot.

Jan 1, 2007

By Korsch RJ

Permian oroclines in eastern Australia must have formed by sliding on a detachment surface. It is identified as a westward- dipping Devonian and Carboniferous sub- duction zone. Major strike-slip faults probably sole out on the same surface. One major fault with an associated serpentinite belt was pulled away from the craton abd bent into loops. Basins formed in the loops.

Jan 1, 1987

By S A. Ravishankar

An overview of flotation reagents used in sulfide and non-sulfide separations, from an industrial perspective, is given in this presentation to highlight the current applications knowledge. The historical and theoretical aspects, which are well documented in numerous books and reviews, will only be highlighted, when necessary, to support the theme of this paper, viz critical analysis of reagent research, design, selection and optimisation with an emphasis on relevance to plant needs and practice. The three types of flotation reagents; viz collectors, frothers and modifiers, used currently in the industry are briefly reviewed; pitfalls and gaps, arising from somewhat arbitrary and reductionistic practice in reagent applications, are identified; and the impact of environmental regulations on current and future reagent development is discussed. The importance of frothers and modifiers, which have received less overall attention than collectors, is highlighted. The need for a rational, holistic (or ætotal systemÆ, ætop-downÆ) approach for reagent design, selection and optimisation is discussed with a view to provide remedies for pitfalls and a platform for filling gaps in reagent chemistry and applications, and to develop reagent technology that can deliver robust solutions to plant needs. Flotation Matrix 100TM, a comprehensive package designed at Cytec to incorporate all the elements of the holistic approach û viz definition of mineralogy and plant needs, rule-based expert system rational reagent selection, and laboratory and plant best practices û will also be described.No paper available for this abstract

Jan 1, 2005

The equipment and general smelter layout for the Mount Isa Copper Smelter were designed by engineers associated with the American Smelting and Refining Company and by Mount Isa staff engineers. The final notice of site was dictated by the need for 'building the smelter near the existing Mount Isa power house to enable full use to be made of the steam from the waste heat boilers. Although much excavation work was necessary before concrete foundations could be poured, the location lends itself to further expansion should it become necessary. The plant is designed and constructed to involve straight line smelting operations which include the following major steps in the overall scheme.1. Charge make-up.2. Roasthig.3. Reverberatory smelting.4. Converting.5. Refining.6. Casting.CHARGE MAKE-UPA system of proportioning bins is provided for preparing the feed mixture for the roasters, Sulphide concentrate containing 24% copper wiil be delivered by belt conveyor from the filter plant to the smelter storage patio. This is a large roofed concrete floor that provides a maximum storage capacity of 8,000 tons of concentrate. Eight proportioning bins located below but adjacent to the south end of the patio will be used for actual charge make-up from which concentrate, ore and flux will be proportioned to the feed belt and conveyed to the roaster feed hoppers. Four of the proportioning bins, 80 tons capacity each, will receive...

Jan 1, 1953

Managing project exposure and risk has numerous viewpoints, including those of engineers, owners, bankers and insurers. Fewer stakeholders are involved in early stages, though with wider risk management techniques and tools available. The costs and sometimes benefits of the occurrence of risk and uncertainty have different contingent outcomes for different stakeholders. For some, the impact of risk outcomes can be mitigated or managed. Some sectors and stakeholders can manage resources project risks more completely, by choice or by externally imposed forces, opportunities and constraints. The core to resources project risk management lies in the identification and analysis of project risks, the examination of project management risk options and alternatives, and the determination of which risks are able to be controlled, managed or laid off to third parties. Without prior risk identification and knowledge, the wildcard of previously unidentified or unconsidered project risk impact has the potential for fatal impact to one or more stakeholders. A range of case studies is reviewed, from exploration projects to development and operating projects, with relevant learning experiences, from the risk management lessons available from the oil and gas sector as well as the mining sector. For many cases, the emphasis focuses on the risk analysis modus operandi and on risk management and risk mitigation techniques that can be and in some cases have recently been implemented for resources project and related debt, capital markets and equity project fundings.

Jan 1, 2003

By Hawes GW

Until 1980 copper, lead and zinc sulphide slurries have been success- fully dewatered for rail and sea transport for many years at the Rose- bery site of North Broken Hill Holdings Ltd. utilizing conventional vacuum disc filtration systems. An upgrading and mill expansion program completed in 1981, which also permitted the introduction of ore streams changes, had resulted in difficulties dewatering copper and lead sulphide slurries to comply with transportable moisture limits. Following an extensive investigation which included both laboratory and pilot plant testing a program was commenced to replace the 2 vacuum disc filtration systems with 2 pressure filtration systems for use on copper and lead slurry dewatering and also to retain other existing vacuum disc filters as standby. An analysis of mill operations performance resulted in slurry storage tanks with a capacity equivalent to 20 hours of normal production being incorporated into the project. The pressure filters are mounted above the concentrate stock- piles. A sealed sloping floor incorporating launders is used to prevent any accidental contamination of the concentrate from pipe leakages, spillage, or from water hoses used for clean up purposes around the area. The project was completed in 1985 and operating performance has been better than that predicted from testwork. Equipment availability at 95076 together with the slurry storage capacity provided results in minimal requirement for the standby vacuum filtration system. Some typical operating results are given.

Jan 1, 1988

A simple method of designing effective fill drainage systems for slurry backfill is presented. By using the proposed method, it is easy to calculate and predict the expected fill drainage rates or to provide additional drainage systems to accelerate the fill drainage. The fundamentals of hydraulic fill and an efficient method to determine its permeability is also explained. The safety in hydraulic filling and bulkhead stability depends totally on achieving effective drainage regime for the fill. This paper provides the basis for creating such efficient drainage regimes. The recent fatal fill bulkhead failures in Australia were the result of poor drainage from the respective fill masses, and therefore it is more important to provide suitable drainage arrangements to continue using hydraulic fill to the benefit of the mining industry. At present there is no procedure available to mine operators to estimate the expected bulkhead pressures to design a satisfactory bulkhead. Some mines are constructing bulkheads after their initial investigations carried out years ago. Other mines have simply borrowed the ideas from the pioneer mines and using in their areas. It is high time that we develop a rational approach to estimate the expected bulkhead pressures for the design of bulkheads. We need a simple model that could be understood and used by the mine operators on a day-to-day basis. This paper also presents a model to predict the bulkhead pressures for designing safe bulkheads.

Jan 1, 2002

This paper comprises a discussion of standards for comparing conditions in mines, of methods for measuring the flow of air, of a comparison of fan performances, and the application of these matters to mines in Western Australia.UNDERGROUND CONDITIONSIn approaching the question of mine ventilation one naturally searches for some standard by which to measure the conditions in a mine.There are several clauses in the Mines Regulation Act defining the conditions which are permitted. The quantity of CO2 is limited to 0·25 by volume, but the estimation of CO2 involves a chemical determination and is not suited for general use in a mine.The maximum temperature is fixed at 87 0 F. (dry bulb), and at 80° (wet bulb). The reading of temperature is easily done in the working places, and this is a very valuable and practical test of the conditions in a mine.The wet bulb temperature is probably the most important factor in determining working conditions. It is obvious that human life cannot be sustained for any appreciable period if the wet bulb temperature is above blood heat (about 97°F.). In places where the wet bulb temperature is 90°F. there is danger of heat stroke.In places where the air in a mine is still, the wet bulb temperature rises to a value close to the rock temperature.Adequate ventilation cools the workings by transference of heat from the rock to the air stream, and the temperature...

Jan 1, 1939

By Hargraves A. J

At 4 MPa gas pres ure and 30°C the sorptive capacity of 10 selected Australian dried coal ranges from 17 to 23 m3/tonne of CH4 and 27 to 36 m3/tonne of CO2. Values for moist coals are about 25 per cent less. At 20°C, sorptive capacities are about 18 per cent greater. No clear relationship between coal rank and sorptive capacity was established.The permeability of coal to gas is dramatically reduced by progressive hydrostatic confinement. No single 'permeability' for a coal can be established because of relative affinities of coals and the gases (Ar, CH4 and CO2) used in experiments. Bedding plane permeability is significantly greater than permeability normal to bedding. Dependent on gas/coal combinations, permeabilities without confinement ranged from 150 to 0.05 millidarcies, and with 8 MPa hydrostatic confinement from 1.0 to 0.001 md. Generally, permeability is increased by reduction in temperature, and reduced by increase in moisture as one changes from Ar through CH4 to CO2. In retrospect, it is preferable to ubject permeability sample to 8 MPa hydrostatic stress only once, because changes occur in the samples and it is difficult to use other gases successively.The initial rate of desorption, DP, does not appear to be a reliable index of proneness to instantaneou outbursts for Australian coals.The methods used for determining quantitative gas-coal properties appear suitable and reliable for testing other coals.

Jan 1, 1985

By Mathews K. E, Willstead T. V C

Initial glory holing was followed by sublevel stoping, with a small proportion of total production mined by shrink and cut-and-fill methods.The sublevel stoping method was successfully modified to cater for altering ground conditions encountered at increasing depths. At present, 94 per cent of production comes from sublevel stopes. The evolution of this mining method and current techniques and practices is outlined. A combined shrink and,sublevel benching method was also developed to increase the mining rate of narrow orebodies over the standard stoping lift of 380 ft. Pillar recovery and stope filling operations have been continuously modified to suit the relIuirements of a rapidly developing mine.INTRODUCTIONOre production commenced at Mount Isa Mines in June 1931. Since then production has increased from approximately 2,500 tons of lead ores per mine day to the current 11,000 tons of sulphide ores per mine day comprising 3,100 tons of lead ore and 7,900 tons of copper ore. The early planners selected sublevel stoping with ring drilling for ore breaking as the most efficient and economical for Mount Isa conditions: This method satisfied requirements for:1. a low cost mining method for the comparatively low grade ores,2. a stoping sequence allowing development, stope preparation, breaking and drawing of ore to proceed concurrently,3. a method of extraction in which broken ore in the stopes would be kept to a minimum because of adverse effects of oxidation on mill metallurgy and...

Jan 1, 1961

By Gregg PEH, Stewart RB

Reclamation field trials commenced in 1985, two years before the granting of the Mining Licence. The level of pasture production on varying soil depths (0 cm, 5 cm, 15 cm and 25 cm) overlying both tailings and oxidised waste overburden was assessed. After seven years, the rate of increase of N, C, P and S, losses of S and P and pH changes were investigated on the plots where grass seed was sown into bare tailings. It was determined that on these plots it would take approximately 24 years to reach the original carbon and nitrogen levels at the mine site.

Jan 1, 1995

In the Massif du Sud ophiolite complex platinum group elements occur predominantly as ultrafine-grained inclusions of rutheniridosmine, laurite and iridosmine. These form primary inclusions less than 5 to 10 um in diameter within chromite in podiform chromitite fringed by dunite, and emplaced subsequent to solidification of the host harzburgite. There is no specific association with trace inclusions of transition metal sulphides. The crystallization of then: Ir. Os and Ru minerals may have occurred concomitant with hromite, perhaps related to more reducing conditions produced by crystallization of the chromite. The Ir, Os. and Ru show pronounced fractionation from Pd, Pt and Rh expected to be also present in the mantle source rock. The dunite shows evidence of metasomatic origin from adjoining harzburgite, and the possibility of selective crystallization of the Ir sub-group elements suggests that Pd, Pt and Rh may have been transported elsewhere.

Jan 1, 1989

This paper traces briefly the reasoning which led to the decision to extend the existing internal shaft at the Rosebery Mine and to instal underground a larger winder (No. 2 Hoist) with capacity to handle all hoisting requirements to a depth of approximately 2,500 ft. on an incline varying from 83" to 55" to the horizontal. It then covers the specifications and selection of the winder, the design of the winch chamber, rope raise and bins, the excavation, support and lining of the chambers and the transport to site and installation of the machinery. Finally a brief description is given of the mechanical and electrical parts of the winder.INTRODUCTIONThe existing shaft winder on the internal inclined shaft situated on No. 8 Level of the Rosebery Mine had reached the limit of its capacity. Hoisting 21-ton capacity skips in balance, with a 13/16-in. diameter rope, it required four layers of rope on the 3 ft. 6 in.drums to reach the loading station under No. 12 level at a depth of 864 ft. (Fig. 1). This entailed heavy wear on the multi-layered rope and heavy stresses on the cheeks of the drums. Continuous working on three shifts was necessary to handle ore, men and materials, and a hold-up from any cause whatever meant an interruption to production on the lower levels of the mine. The maximum output of the hoist whilst hoisting from the No. 12 level loading station was 50 tons per hour. It was obvious therefore, that a larger winder was required if output was to be maintained.Diamond drilling had proved the existence of sufficient ore at depths well below No. 12 to warrant exploitation and a decision had to be reached on the means of hoisting this ore.

Jan 1, 1958