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The most frequently applied project evaluation criteria are the Net Present Value, the Internal Rate of Return, and the Pay Back Period. Unfortunately, the Internal Rate of Return and the Pay Back Period, if interpreted according to their traditional definitions, easily produce results which contradict the Net Present Value. The reasons for the Net Present Value incompatibility of the two criteria are that these criteria have not been adequately defined and so new, complete definitions are recommended. The concept of the True Rate of Return is proposed as a substitute for the interest rate that the Internal Rate of Return has been believed to represent. The redefined Internal Rate of Return and Pay Back Period criteria, as well as the True Rate of Return are all Net Present Value compatible and can, therefore, serve as independent project evaluation criteria. Also, the correct approach to the evaluation of mutually exclusive alternatives and the status of the often neglected borrowing projects in project evaluation are found to be quite important.

Jan 1, 1991

The first stage in the refining of lead blast furnace bullion is the removal of copper and is normally done as a batch process in open kettles.A number of attempts have been made to establish continuous processes operating in reverberatory type furnaces, but the growth of accretions in the furnace has always presented a serious problem.The Broken Hill Associated Smelters Pty. Ltd. (RH.A.S.) at Port Pirie have overcome this problem by constructing a reverberatory furnace and ancillary cooling system of a novel design.Although a number of mechanical and metallurgical problems still remain to be solved, it is claimed that a workable continuous copper drossing process has been established.INTRODUCTIONLead bullion emerging from the blast furnaces at Port Pirie is at a temperature of about 1100°C and contains, among other impurities, about 1 per cent of copper.The first stage of refining is the removal of this copper by a drossing operation and in the past this was done, batchwise, by allowing the bullion to cool back in 100 ton kettles.At about 700°C, copper dross begins to separate as a crust on the surface of the kettle and this separation was allowed to continue down to about 450°C when the crust was brokenup and skimmed manually using a long handled, balanced, spoon. The dross was then smelted in a reverberatory furnace to produce a saleable copper matte and separate out lead bullion, which was returned to the drossing kettles. The copper content of the drossed bullion was then further reduced by a sulphur treatment process.The drossing operations involved hard manual labour under arduous conditions and the pouring of hot bullion into...

Jan 1, 1963

The Kidston Gold Mine i s located in northern Queensland some 2 Ukm northwest of Townsville. I t is the largest gold mine in Australia and produces approximately 25 percent of Australia's gold and one percent of the world's gold.Initially, traditional mining industry management approaches resulted in the rejection of what appeared to be an attractive low-grade deposit as an uneconomic proposition. Critical re-examination revealed that the rejection was based upon managerial inflexibility and traditionally over-conservative methods of mine evaluation, construction, operation, projected commodity prices and contingency factors. "Ihus it tradition was to prevail, the Kidston prospect would never have been realised.

Jan 1, 1985

By Rogers J. F, Bird A. C

The establishment of vegetation for dust control on tailings dams at The Zinc Corporation, Limited and New Broken Hill Consolidated Limited (ZC/NBHC) is outlined. The evolution of practices and techniques employed is described and the problems encountered are discussed.

Jan 1, 1983

By Secis R

Northparkes Mines is located 27 km northwest of Parkes in central west New South Wales. The E26 deposit is the largest of all defined to-date in the area, containing over 50 per cent of Northparkes' resources and reserves. The combination of orebody geometry and geotechnical characteristics provide ideal conditions for the block cave method of mining of the first lift of the E26 deposit. Although this method of underground mining has been used for decades in other countries, E26 is the first example of its use in Australia. Copper-gold mineralisation at Northparkes is related to quartz monzonite porphyries that have intruded coeval Ordovician volcanics. Sulphide mineralisation (dominantly bornite and chalcopyrite) occurs principally within quartz veins associated with the porphyry intrusions. Veining and mineralisation occurs within both the porphyries and volcanics. To optimise the recovery of a resource within a block cave mine, a good understanding of the rock mass and ore flow characteristics is essential. These have a significant effect on the fragmentation size and range and on parameters such as the drawzone spacing, height of interaction zone and dilution entry. As ore within a cave column moves and mixes when material is extracted from below and beside it, the total reserve extracted is unique for a particular draw strategy/production history. Until recently, reserve estimates and production schedules were based on the application of a uniform dilution factor to the tabulated resource. However, given the complicated flow and mixing within the cave column, such a simplistic approach can produce erroneous results. Northparkes Mines purchased and sponsored the upgrading of a commercially available software package specifically developed for modelling removal of resource blocks under caving conditions. This Gemcom software package, PC-BC, has been used for mine planning, reserve reporting and production scheduling. Blocks within it resource block model are converted into a mineable reserve taking into account mixing and dilution during depletion of drawcolumns (by horizontal and vertical redistribution of insitu tonnage and metal). As PC-BC gives individual drawpoint reserves, it can be used during the planning stage to determine whether or not a drawpoint is economic. For E26 Lift I an optimal layout containing 130 drawpoints has been developed based on a nominal 14 m x 14 m drawpoint spacing and a 0.8 per cent equivalent copper (eCu) cut-off grade. Annual production schedules for the depletion of material above this layout are based on: I higher tonnage removal from higher grade drawpoints in the early years of production; and 2 drawpoint production rates proportional to the total tonnage of orr remaining in the drawcolumn up to the optimal height of draw tier subsequent years. The reserve is tabulated from the combined annual production schedules over the life of the lift. The pre-mining reserve for Lift I of E26 is 26.6 Mt at 1.43 per cent copper and 0.39 g/t gold, of which approximately 23.4 Mt will be extracted under caving conditions. By scaling down the annual production schedule, weekly drawpoint tonnage calls can be given for short-tern production scheduling.

Jan 1, 1997

The Mineral Research Centre is located at Perth in the suburb of Waterford, adjacent to Curtin University of Technology and within a Technology Precinct which includes Technology Park. The Centre, constructed by the WA Government at a cost of $16 million and occupied in late-1994, comprises three separate buildings housing the CSIRO Division of Minerals, Curtin University's School of Chemical Engineering and the Mineral Processing Laboratory of the Department of Minerals and Energy. The emphasis on R&D at the Centre is primarily in hydrometallurgy as a result of staff expertise, extensive bench and pilot-scale facilities, track record with the industry and, very importantly, two of the occupiers being partners in the AJ Parker Cooperative Research Centre for Hydrometallurgy. Staff in the Centre have a long involvement with R&D agencies such as AMIRA and MERIWA. The hydrometallurgical focus of the Centre complements that of pyrometallurgy at the GK Williams CRC in Melbourne and the Centre for Mining Technology and Equipment in Brisbane, to ensure that the mineral processing industry in Australia has access to a complete range of expertise and facilities to serve its needs.

Jan 1, 1996

By Jones S. C

A specific arc can be cut in hard rock formations using a specially designed multistep piloI bit, variable sized drift joint and long heavy control rod as a stabilizer rather than the Vaughan Thompson "Lifting" arc cuttingequipment. The rate of deflection can be controlled after an initial determination of behaviour in any particular rock type.

Jan 1, 1966

The methods and equipment developed and used at the Mount Isa mine to obtain samples of uncemented hydraulic fill lying above major horizontal pillars are described. From existing underground development, samples were obtained at distances up to 50 m, of which up to 30 m was through hard rock. Samples were used to predict fill behaviour as it affected mining security during recovery of the pillars.Specialized equipment for difficult sampling situations and also other sampling exercises are described.

Jan 1, 1976

By G W. Wilson

Rehabilitation of tailings impoundments is one of the most challenging aspects in mine closure, as not only does the potential for producing leachate pose a challenge to the rehabilitation designer, but also other aspects such as stability and settlement must be considered. The water balance of a tailings impoundment is unique in the sense that it usually hosts a pond that in turn causes a phreatic surface in the impoundment. The position of the phreatic surface defines the saturated and unsaturated zones in the impoundment, which of course varies spatially and temporally. Predictive modelling for this hydrologic system becomes difficult, as numerical models capable of analysing the combined saturated/unsaturated zones are not adequately refined to accurately solve the flux boundary problem for infiltration at the surface of the tailings. This paper describes the development of a flux boundary model that enabled accurate modelling of the unsaturated zone in the tailings impoundment at Kidston Gold Mine, Queensland, Australia. The technique made it possible to accurately predict the spatial variation of infiltration to the tailings as a result of the presence of the phreatic table. The conceptual model defines the boundaries for the Kidston tailings dam, based on a generalised cross-section through the dam. Firstly the shape of the top boundary is described as a function of the particle size of the tailings associated with particle segregation that occurs due to the hydraulic placement of tailings. This function was rigorously tested using in situ infiltration tests and laboratory materials testing. Defining the bottom boundary of the section, ie the phreatic level was done by physically monitoring the phreatic level with 42 piezometers strategically located on the tailings dam. Analysis of this data led to the fixing of this bottom boundary through the introduction of a function that mimicked the shape of the phreatic level. The material properties of the surface layer, specifically the saturated permeability that would govern the infiltration/exfiltration rates of moisture through the top boundary layer were defined on the basis of extensive field and laboratory material testing. A function for the saturated permeability was developed on the principles of particle segregation. Finally a procedure is described for the application of the new conceptual model for the evaluation of the spatial infiltration to a tailings surface using the rigorous one dimensional surface flux boundary code SoilCover. The solution of the calculation described above was then used as the top boundary condition in conventional three dimensional saturated/ unsaturated flow codes, and verified against measured seepage rates from the tailings impoundment under drains. It was concluded that the developed spatial flux function was in deed a most rigorous representation of reality. Although the methodology presented here was applied to the Kidston tailings impoundment only, the authors believe that the principles are directly transferable to any tailings impoundment that has a varying phreatic zone thickness due to the presence of a pond. The mathematical equations describing the typical cross-section would be site specific, but the approach and ultimate solution methodology would be directly transferable.

Jan 1, 2003

By Mowat BA, Leach TM

The Morobe Goldfield in Papua New Guinea has produced in excess of 120 tonnes of gold from predominantly alluvial sources. Hardrock gold has been mined historically at Wau (23t) and Edie Creek (40 and geological resources have been established at Hidden Valley (780, Kerimenge (360 and Hamata (30t). Numerous prospects have also been identified within the goldfield. Primary gold mineralisation is related to intrusion of dacitic to andesitic Edie Porphyry (2.4 to 3.8 Ma), often with associated diatremes, which have intruded along major regional structures. A full spectrum of deposit styles from proximal (mesothermal) to distal (epithermal) to the porphyry source is recognised. The transition between the proximal and distal styles is recognised both between deposits and within an individual deposit. The deposit styles should be termed porphyry related gold mineralisation rather than given the generic epithermal and mesothermal labels.

Jan 1, 1995

This paper reviews an analytical model that defines the relationship of weight-weighted or volume-weighted average haulage distance with bench height and other related parameters; presents the optimum bench heights under various conditions in terms of minimising the weighted average haulage distance of ore/waste during the full lifetime of an open pit mine, and discusses the truck-haulage economics of mining open pits by high bench, specifically by 15 m, 18 m and 20 m bench through marginal analysis. The observations from these analyses disprove the previous conflicting assertions that the overall weighted average haulage distance either increases or decreases monotonically with bench height. It is concluded that: (i) an increase of bench height to 15 m from 12 m is likely to be viable under the general conditions of large-scale deep open-pit mines; (ii) the reduction of haulage cost is limited with the case of increasing bench height to 18 m from 15 m; and (iii) the feasibility of mining open pits by 20 m benches is generally questionable. These conclusions may serve as a useful guide for those who advocate open pit mining by high benches.

Jan 1, 1995

The process of mine planning at AnacondaÆs nickel and cobalt laterite operation located at Murrin-Murrin in the Northern Goldfields of Western Australia is carried out in a number of stages in an endeavour to determine the optimum mining sequence in order to maintain supply to the 4 Mtpa plant subject to operational constraints. The 296 Mt @ 0.99 per cent Ni and 0.064 per cent Co ore reserve (based on a cut-off grade of 0.8 per cent Nickel) is spread geographically widely over three regions comprising Murrin North (15 km + 15 km), Murrin South (7 km + 16 km) and Murrin East (5 km + 13 km) (refer Figure 2). The nickel mineralisation has been broken up further within each region into 18 separate manageable block models representing the 32 different deposits. The models are based initially on the 50 m + 50 m resource drilling for long term planning and updated with 12.5 m + 12.5 m grade control drilling prior to mining. Each of the individual block models are optimised using Whittle 4X software to provide the shells used as the basis for each of the 126 pit designs. The pits are generally nested with starter pits targeting the higher value shells with subsequent pits incrementally cutting back ultimately to the life of mine shells. A series of steps including: the coding of the individual pits into the block model; data export; TCL scripting (to calculate additional economic and tracking information); importing; and re-blocking; completes the transfer of the data into a MineMax (mixed integer linear programming schedule optimising software) model. Using MineMax and the application of operational and production constraints allows the optimisation, assessment and economic evaluation of alternative scenarios in order to determine the optimum life of mine plan and schedule.

Jan 1, 2003

By P J. Hawke

Within the Mesa J pisolitic iron ore mine, clay is distributed unevenly throughout the deposit and often forms as pods and channels less than 10 m in width. The mine resource model, based on 50 m + 50 m spaced drilling, has been unable to accurately delineate the clay for scheduling and mining purposes. Therefore, alternative methods for mapping the clay-contaminated ore were sought. Predictive clay mapping will significantly reduce production costs at Mesa J through improved selective mining practices, more accurate scheduling, reduced wear on mining equipment and minimising the need for maintenance of haul roads and ramps. High lateral resolution electromagnetic (EM) surveying using a Geonics EM31 instrument can quickly and inexpensively delineate the extent and orientation of the clays. Preliminary results indicate that a block model based on the interpretation of the EM data can accurately predict the distribution, volume and tonnage of ore and waste material. While rip-line mapping, infill drilling, ground penetrating radar and gravity can delineate clay with various degrees of success, none were as efficient as EM31.

Jan 1, 2002

By R M. Tosdal

Intrusion related ore deposits are widespread in the Central Peruvian Cordillera Occidental and adjacent high-plains to the east and include porphyry Cu (-Au, Mo; eg Morococha), Cu-Zn (Pb-) skarn (eg Antamina), Pb-Zn-Ag carbonate hosted vein deposits (eg Uchucchacua), cordilleran base metal (eg Cerro de Pasco) and epithermal deposits (eg Quicay). Geochronology and petrochemistry of intrusive rocks associated with mineralisation reveals that: most ore deposits were emplaced in the late middle to early late Miocene; and are related to calc-alkaline granodioritic igneous rocks with Sm/Yb >3, Sr/Y >40 and Y/Th <1.5. The largest deposits, including Antamina, Colquijirca and Cerro de Pasco are associated with intrusions with Sm/Yb >4 and which, as indicated by Pb and Sr isotopic data, have undergone relatively moderate middle and upper crustal assimilation. The chemical signature of igneous rocks related to mineralisation likely reflects melt equilibration in the lower crust at a crustal thickness of more than 45 km. Mineralisation was favoured by a change from normal to flat subduction, resulting in increased compression, crustal thickening and subsequent cessation of the magmatic arc. Premiocene ore deposits do occur, but are restricted to a broad E-W corridor between Uchucchacua and Milpo and probably were controlled by a crustal-scale structural discontinuity across the strike of theAndes. Not all igneous rocks with high Sm/Yb and Sr/Y ratios did acquire their trace element signature in the lower crust. Petrographic and Pb isotopic evidence indicates shallow crustal hornblende fractionation or assimilation of pelitic material as possible causes for high trace element ratios in some rocks. These are apparently unrelated to polymetallic mineralisation. Thus, critical evaluation of petrographic and geochemical data in metallogenetic studies, especially in areas with complex basement architecture such as Central Peru is critical and should not be reduced to simple trace element ratios. An EXTENDED ABSTRACT is available for download. A full-length paper was not prepared for this presentation.

Jan 1, 2008

By Gelding P. W, van Herk H

The Cattlegrid copper orebody is a concealed thin, tabular sheet localised near an unconformity between two sub- horizontal Precambrian sandstone units. It extends over an area of 1,000 m by 500 m within a wider area of mineral- isation. The ore consists of complex sulphides and is neither massive nor disseminated. It consists mainly of irregular fillings in pore space developed between clasts in sandstone breccias, in fault zones and joints and along stratification planes. Its habit is unique amongst Australian base metal ore bodies. During the pre-development period ore was outlined by percussion holes but in studies leading to investment decisions ore reserves were manually calculated from intercepts and assay results of cores from diamond drill holes. The methods used for estimating reserves in the pre-production and production periods are described and evaluated by comparing mine production during the first five ore production years with the forecasts made.

Jan 1, 1979

By Thimons E. D, Derick R. L

In early June, 1986, a fire of unknown origin erupted in one of several main entries of a coal mine located in southwestern Colorado. The fire was aggressively fought with high-expansion foam. As a result of several strategic decisions, the fire appeared to be held in-check. Unfortunately, sufficient quantities of foam were unavailable and as soon as the foam was depleted, control of the fire was lost. To prevent further damage and the possibility of an explosion, all entries to the mine were sealed.

Jan 1, 1988

By Mack JF

A metallurgical evaluation of the Argyle Kimberlite No. 1 (AKl) orebody has been carried out using a 40 tonnes per hour pilot plant. Pilot plant results, together with laboratory investigations provided a basis for the determination of process design criteria for a major treatment plant. Various sampling programmes were carried out to determine specific aspects in the evaluation of diamond liberation. Treatment parameters such as the top crush size, recrush size and the bottom cut-off size were investigated. A description of the pilot plant is given together with the design criteria for a 3 million tonne per annum treatment plant

Jan 1, 1984

By Taylor G. R

Fluid inclusions within quartz accompanying several stages of mineralization at Hillgrove are de cri bed. Pyrite-arsenopyrite-gold mineralization shows uncorrected filling temperatures of between 195°C and 250°C. Later stibnite-antimonide-siderite and stibnite-calcite mineralization has filling temperatures between 100°C and 195°C and a bimodal distribution of temperature peaking at 175°C and 145°C, possibly correlated with the two stages of mineral deposition. Secondary included fluids within an earlier generation of deformed quartz veins are derived from hydrothermal fluids contemporary with the main-stage mineralizing fluids described above. Primary inclusions within latestage, cross-cutting quartz-chlorite-calcite veins show these veins to be unrelated to the mineralization.A correlation between filling temperatures and salinities indicates that mineralization accompanied the mixing of a hot, partly saline, fluid (at 5 per cent equivalent NaCI) with cooler, less-saline meteoric or connate water. Over a vertical distance of 200 m, equivalent mineral assemblages were deposited over the same temperature range. Studies on single crystals suggest rapid deposition with fluid temperature dropping 45°C during their growth. Ore mineragraphy and the fluid inclusion data suggest that the metals were transported by sulphur-poor, relatively low-temperature, low-salinity solutions, possibly as carbonate complexe , and that a vertical zonation of the W-Sb-Au mineralization is compatible with the mixing of this hydrothermal fluid with cooler groundwater. Regional metallogenic and geological considerations suggest that ore metals are derived from the remobilization of deep-seated stratabound deposits or from metamorphic fluids expelled from deepseated volcanoclastic sediments.

Jan 1, 1984

By Simes C. D

Results of experiments for dissolution of zinc by aeration of sphalerite in glass, copper, galvanised iron and stainless steel vessels of identical dimensions are reported. It was found that in a glass container dissolution of zinc reaches a maximum after 30 min aeration and then decreases whereas in the copper container dissolution remains nearly unchanged up to 30 min aeration and then increases. In the cases of galvanised iron and stainless steel containers zinc dissolution increased steadily with time up to 60 min aeration. It is suggested that the different rates of dissolution in glass and copper containers may be due to electro-chemical actions.

Jan 1, 1984

In the latter half of 1973 a contract was given to Roberts-Holland Joint Venture for the stripping of the southern exhaust airway.The contract was carried out efficiently and relatively cheaply. It was completed 2 months ahead of schedule.The sinking stage and in particular the drilling jumbo proved to be a success and helped produce an above schedule stripping rate.No major problems were encountered with ground conditions or ventilation. A survey of the shaft showed the diameter to be within the set limits of 5·49 m ± 0·15 m.

Jan 1, 1974