Search Documents

By Bertrand D. L

Although investigations into treatment of Mt. Keast (a hill of tailing) began as early as 1940, it was not until the Imperial Smelting Furnace (I.S.F.) was installed at Cockle Creek by Sulphide Corporation that its treatment became eooncmically feasible. Early investigations and milling operations which began in mid-1976, with their developments are described. Metal recovery involves sulphide flotation, oxide flotation and gravity techniques.

Jan 1, 1983

By J M. Santini, I C. A Streimann

Arsenic is a naturally occurring element in the environment and is toxic to biological systems. Arsenic can exist in the environment in the following redox states, arsenite (+3), arsenate (+5), arsine (-3) and elemental (0) (Oremland and Stolz, 2003). Arsenic is released into the environment by natural and anthropogenic activities such as rock weathering and ore mining, respectively. Such activities result in the release of the soluble inorganic arsenic oxyanions, arsenite and arsenate. The redox state of arsenic in the environment is influenced by a number of factors including, redox potential, pH and microorganisms (Irgolic et al, 1977; Genc-Fuhrman, Tjell and McConchie, 2004). Recently, the ability of microorganisms to influence the cycling of arsenic in the environment has attracted much attention. Microorganisms that can oxidise arsenite and reduce arsenate have been isolated and the mechanisms of arsenic metabolism are currently being investigated. The differences between arsenic resistance and arsenic metabolism are being revealed. The role of microorganisms in the release of arsenic into the environment is of particular importance with arsenic contamination of waters causing the death of millions of people worldwide, due to the consumption of contaminated water (Vasken Aposhian et al, 2003; Islam et al, 2004).

Jan 1, 2004

The Honourable Neil Pickard has been in Parliament for something like sixteen years. He was Minister for Education in 1976, served as Shadow Minister in a number of portfolio areas while his party was in opposition, and since their return to power in 1988, he has been the Minister for Minerals and Energy. Ladies and gentlemen, I ask you to welcome the Minister, the Honourable Neil Pickard.

Jan 1, 1989

By S McKnight, L Koroznikova, S Hall

Mineral sands represent an important new resource being developed in the Murray Basin, including parts of western Victoria. This paper will outline a simple methodology for mineral sands characterisation, developed as part of the AMIRA-managed project P777 æThe Development of Heavy Suspension Techniques for High Density Separations (Replacement of ClericiÆs Solution)Æ. This project is currently sponsored by three multinational mining companies (De Beers Consolidated Mines, Iluka Resources Limited and Rio Tinto Limited) and is developing an innovative laboratory mineral characterisation procedure that will allow the replacement of the currently employed highly-toxic chemicals.Mineral sand resources almost always contain more than one valuable (and relatively heavy) mineral. Titanium minerals are found with a large range of titanium contents, giving rise to density variation and often subjective mineralogical descriptions. Companies tend to rely on laboratory heavy liquid separation in the evaluation of samples arising from exploration, mining or metallurgical processes. Unfortunately, there are only a limited number of high density (æheavyÆ) liquids and these tend to be more toxic as their density increases. Low-toxicity inorganic solutions, based on tungsten compounds, have been developed that can be utilised at relative densities (RD) up to 3.0. However, beyond this value currently only organic liquids can be used. Diiodomethane (methylene iodide) having a relative density of 3.31 is commonly used; however, this presents significant health and safety hazards. Mixtures of thallium formate and thallium malonate were found in the early 1900s by Clerici to provide liquids having specific gravities between 4.0 and 5.0, hence æClericiÆs solutionÆ. For the characterisation of the heavy components of mineral sand deposits (eg anatase sg 3.9, rutile sg 4.2, ilmenite sg 4.4 û 4.7 and zircon sg 4.6 û 4.8) there is currently no heavy liquid alternative to ClericiÆs solution. ClericiÆs solution is highly toxic and testing is now conducted by very few laboratories worldwide with costs reflecting the chemical costs (though extensive efforts are made to recover and reuse the liquid, plus the requirement of its removal from the mineral samples), the infrastructure costs and health and safety regimes (eg blood testing of exposed staff, inventory management). A simple laboratory technique of density fractionation is being developed, employing suspensions of fine tungsten carbide particles in lithium heteropolytungstate (LST) solutions, that can replace ClericiÆs solution in the evaluation of fine mineral sands samples (eg -250+150 microns). The developing methodology that can achieve low-cost, low-toxic separations at relative densities above 5.0 will be outlined and the comparison of results with ClericiÆs solution presented.

Jan 1, 2006

Commencing early in 1975, Saraji Mine in Central Queensland will product 3.56 million tonnes in the first year and then 4.57 million tonnes annually for the Japanese and European markets. The coal will be mined from a series of pits separated by corridors which will be mined at a later date. Through the corridors flow regional drainage structures and power lines. Relatively easy areas will be mined initially to assist the cash flow and to minimise start-up difficulties. For ease of scheduling it is desirable that pit lengths exceed 3 000m but this will not always be possible. Ramps were designed using 1 200m spacing and 5 per cent gradient as optimum but modified to suit specific conditions. One ramp will be in the end of each pit nearest the ROM hopper and stripping will commence from this end. The advantages so gained are shorter haul distances, immediate access to the coal after stripping and simplification of the excavation of the first ramp. Haul road location was defined by the minimum present value of construction plus operating costs. Road design utilised airfield construction charts because wheel loads were outside the range of available road design manuals. Low load resistance characteristics of

Jan 1, 1974

By Sullivan T. D

The Lochiel Deposit is a large occurrence of soft brown coal about 130 km north of Adelaide. The deposit is undergoing continued evalua- tion of which an integral part was the excavation of a Trial Pit. The Trial Pit was deemed necessary to demonstrate that the deposit may be mined efficiently by draglines and the particular emphasis was on the geotech- nical and hydrogeological aspects of what may be termed difficult mining conditions. The dewatering of the Trial Pit was commenced in September 1986. After considerable planning and excavation was successfully completed in May 1987. The project is unique in terms of its size and technical com- plexity. It entailed excavation of over 230,000 m3 of overburden to an overall depth of 36 metres. The excavated material was utilised for experimentation in a Trial Dump. Included in the Trial Pit design were several novel techniques includ- ing depressurisation drains to improve mining conditions in soil strength materials and the use of vacuum to aid dewatering.

Jan 1, 1988

By A E. C Peres, J D. Donda

With the aim of producing a concentrate with silica content lower than 1.0 per cent SiO2, laboratory experiments were conducted using a sample of the final concentrate from the Samarco iron ore flotation plant (mechanical cells circuit). The sample contained 3.08 per cent SiO2. This sample was split in two size ranges: >105 ¦m and 2. The >105 ¦m fraction, representing 23.3 per cent of the initial mass, was re-ground to 100 per cent 2 content were obtained. Considering the mass proportions of the two size fractions, the calculated silica content of the combined concentrate was 0.68 per cent, which is considered satisfactory for the objective of this investigation. The major reasons for the higher (0.90 per cent) silica content of the concentrate from the coarse fraction are the presence of silicon in the lattice of goethite and the presence of residual silicious slimes in the pores of porous hematite.

Jan 1, 2005

Minerals similar to the indicator minerals used in kimberlite exploration are widespread in the Cainozoic alkali basalts of north Queensland. Magnesian ilmenite, pyrope garnet, chrome diopside, chrome spinel, phlogopite and zircon were recovered from most of the localities examined. The occurrence of these minerals creates significant problems when indicator minerals are used for diamond exploration. Basaltic ilmenite is visually similar to kimberlitic ilmenite. Untransported grains are rounded and have pitted surfaces. Basaltic ilmenite grains often have exsolution lamellae and reaction rims of titanomagnetite. Although kimberlitic and basaltic ilmenite are chemically similar, basaltic ilmenite grains are usually less magnesian than their kimberlitic counterparts. Basaltic ilmenite contains very little Cr203. Pyrope garnet occurs as megacrysts and in a variety of xenolith lithologies in north Queensland basalts. Most basaltic garnet is Cr-poor and comparable with kimberlitic garnet from Groups 3 and 5 of Dawson and Stephens (1975). Wine-purple, chrome-bearing garnet is common in kimberlite but rare in basalt.

Jan 1, 1984

By J P. Jona, S R. Poulson, C H. Gammons

Spatial and temporal trends in water chemistry in the acidic (pH ~2.6) Berkeley Pit lake, Butte, Montana, can be explained by a newly developed iron-cycling model. Key elements of this model include: subaqueous oxidation of pyrite and other metal sulfide minerals by dissolved ferric iron, producing Fe2+ as a by-product; re-oxidation of Fe2+ to Fe3+ in the epilimnion, assisted by A ferrooxidans; precipitation of secondary ferric minerals (schwertmannite, jarosite) in the epilimnion; and replenishment of the Fe3+ concentration of the hypolimnion by seasonal lake turnover and gravitational settling and re-dissolution of ferric precipitates. Pyrite oxidation rates were approximated by using published rate laws and information on the Fe3+ concentrations (~5 mmol) of the Berkeley Pit as a function of depth. The results indicate that subaqueous pyrite oxidation may be releasing as little as 6 + 106 or as much as 600 + 106 moles of H+ each year to the lake; the large range reflects uncertainty in calculation of the specific surface area of pyrite in the flooded pit. This helps to explain the observation that the pH and metal concentrations of the Berkeley Pit have not improved with time since mine flooding began in 1983. Vertical cycling of ferric iron in the pit lake permits pyrite oxidation to continue in deep water, even in the absence of dissolved oxygen. This hypothesis is supported by preliminary stable-isotope data.

Jan 1, 2003

By A R. Gerson

Pyrite, chalcopyrite and mixed sulfide samples, with and without quartz (5 wt per cent S), were subjected to column leaching to mimic natural weathering (cyclic wetting and drying). The pH and SO4 2- concentration of the column effluent was measured over a period of 78 weeks. The pH of the pyrite-only containing column rapidly decreased and remained under pH 4 after 12 weeks of leaching. The pH calculated assuming a 2:1 ratio of H + generation and SO4 2- release was slightly lower than that measured. However, a reduced H + to SO4 2- ratio of 1:1 is predicted if the release of aqueous Fe 2+ is assumed rather than the formation of Fe(OH)3. The pH of all columns containing only chalcopyrite or a 1:1 mole ratio of chalcopyrite and pyrite remained at or above pH 5 for the duration of the leach even though significant SO4 2- was released. One column containing the same initial surface area of pyrite and chalcopyrite but only approximately 0.10 the mole fraction of S as chalcopyrite maintained a pH of greater than five for the first 56 weeks of leaching and showed sporadic acid production thereafter. It is well established that oxidation of chalcopyrite results in a surface layer of polysulfide material. It is the formation of polysulfide that neutralises the acid produced on S oxidation through to SO4 2-. Hence where polysulfide formation occurs extensively on chalcopyrite the resulting solution pH decrease will be less than that based on SO4 2- production. The oxidation rate of sulfide minerals increases with decreasing pH and hence acid formation via sulfide oxidation to SO4 2- causes the subsequent oxidation rate to increase. The addition of calcite (CaCO3)to a pyrite column not only neutralises the acid formed but also decreases the rate of release of sulfate to solution by a factor of approximately 30. This is also the case where chalcopyrite is present mixed with pyrite. The rate of release of sulfate from the mixed sulfide mineral columns (after the initial release of the sulfate already present) was between one-fifth and one-tenth that in the absence of chalcopyrite.

Jan 1, 2003

By Nilmani M, Malone J, Taylor RN, Floyd JM, Gray NB, Fountain CR

CSIRO staff and University of Melbourne staff and students collaborated in an intense programme of investigation on the CSIRO 50 kg Sirosmelt pilot plant.In two days of operation copper matte converting and zinc slag fuming trials were performed, whilst investigation of the process dynamics and process chemistry was carried out. This exercise is an example of the techniques available for process evaluation, which can rapidly advance the understanding of pyrometallurgical operations by use of a team approach covering a wide range of experiments.

Jan 1, 1985

By M G. Hawtin

The Rocklands copper project is a multilode, iron oxide copper gold (IOCG) + cobalt mineralised system located in the Eastern Fold Belt of the Mt Isa Inlier. Rocklands orebodies cover a range of oxidation zones from oxide, to supergene and primary. Each zone contains different economic minerals of interest. The supergene zones contain significant amounts of native copper and the primary zones are dominated by chalcopyrite. The juxtaposition of these oxidation zones and minerals make it an interesting mineralisation system.Current mine production sampling collects a nominal 3 kg drill chip sample from a 1 m composite, RC drilled interval. A site specific sampling procedure was needed to quantify introduced sampling error when taking grade control samples. Due to oxidation and mineralogical variation in ore zones we needed to develop a practically achievable sampling methodology that ensures minimum sample mass and sample error requirements are achieved to ensure ‘representative’ samples for all ore types.These issues are planned to be resolved using an experimentally derived method using principles from the Theory of Sampling. This method uses an experimental calibration of parameters K and a for Gy’s fundamental sampling error formula using a modified 32-piece sampling tree experiment to compile sampling nomograms for the primary and supergene ores. This paper demonstrates the process undertaken to improve the quality and standard of existing mine production sampling procedures at the Rocklands operation.Results indicate a sample mass of 5 kg for RC drill chips is appropriate for both the primary and supergene ores to achieve a fundamental error below 0.01 or 10 per cent relative error, which is slightly higher than the existing sampling protocol of 3 kg. An additional experiment is recommended to compile sampling nomograms using the exact laboratory comminution and mass reduction stage information to assess the use of a 3 kg sample. The results also indicate no changes are required to existing sampling protocols when sampling RC drill chips in primary and supergene ore at Rocklands.Sampling nomogram results also provide a useful ‘rule of thumb’ reference guide to other mining and exploration companies to help understand minimum sample mass requirements and sampling error during exploration drilling, resource evaluation and mining production stages when targeting similar IOCG mineralised systems. A sample mass of 5 kg is recommended when RC drilling similarly IOCG mineralised prospects and deposits in the Cloncurry area. Future test work is recommended when additional open cuts become active as well as including resource drilling drill core to provide better three-dimensional sample representivity.CITATION:Hawtin, M G, 2017. Minimum sample mass from reverse circulation grade control drilling at Rocklands copper project, in Proceedings Tenth International Mining Geology Conference 2017, pp 337–346 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 20, 2017

By Harada M, Katsuta S

In order to shorten the time taken to com- plete shield tunnel excavation work, reduce labor costs and to improve safety, an automatic assembly robot system for the segments used has been developed. It was developed in the three years from 1982 until 1984. An experimental assembly system for concrete segments (outer diameter:.3.55m) has been thoroughly tested to confirm that these segments can be fed and assembled automatically (Fig. 1, 2). This report outlines the robot which will be put into practical use in the near future. The robot senses the position of already assem- bled segments by means of position sensors, controls the position and the attitude of the segment to be assembled to allow bolt hole positioning, and completes the automatic assembly of a se;nent by automatically tighten- ing bolts. The six segments which make up one ring can be automatically assembled in about 25 minutes.

Jan 1, 1987

The Louvicourt Mine is located 600 km north of Montreal in the province of Quebec, Canada. The owners are Aur Resources Inc (30 per cent), mine operator, Novicourt Inc (45 per cent) and Teck Corporation (25 per cent). It is an underground mechanised operation with a total work of 290 employees. Since 1994, the mine production is 11 120 456 tonnes at 3.60 per cent Cu, 1.51 per cent Zn, 26.07 g/mt Ag and 0.96 g/mt Au. At a rate of 4300 tonnes per day 365 days per year. Louvicourt Mine is a copper operation since 80 per cent of our revenue is from copper. The mining method is transverse blasthole stoping, mining primary and secondary stopes. The stope dimension is 15 m wide by 30 m high and 10 m to 80 m long for primary stopes. The secondary stopes are limited at 25 m long. Close to 100 000 metres of production drilling per year is required for 115 mm and 165 mm diameter holes. Backfilling is done with pastefill from the start of the operation in 1994. The binder ratio is 4.5 per cent in primary and two per cent in secondary stopes (80 per cent slag and 20 per cent cement). Development requirement to maintain production was 5000 m from 1996, 1997, 1998, 4200 in year 1999, 3200 m in year 2000 and an average of 2500 m in 2001 and 2002. Typical 6.0 m + 4.3 m, 4.9 m + 4.6 m and 4.5 m + 4.6 m headings. In addition more than 1500 metres of development will be excavated in/under paste backfill before 2004. A total of 60 mobile units are required among which for lateral development (three two boom electric hydraulic jumbos û AC Boomer 322, four electric hydraulic roofbolters û Maclean MEM-928 and AC Boltec PL350 and two anfo loader truck û Dux) and for production (four ITH drills û MTI, one electric hydraulic longhole drill û AC Simba H1357, eight LHD û AC-Wagner ST8-B, 1 LHD ST-1000 and two haulage 30 tons trucks û Dux. The 1 January 2002 reserves are 4.1 M tonnes grading 3.1 per cent Cu and 1.9 per cent Zn at an NSR (Net Smelter Revenue) of $US 53. The operating mining costs were at an average of $US 23 from 1996 to year 2000 and $US 25 in 2001. The operating costs ranged from a low of $US 0.41 per pound of copper in 1999 to a high of $US 0.49 per pound of copper in 1998.

Jan 1, 2002

There has been much debate in the past as to whether the Australasian Institute of Mining and Meteorology ("AusIMM") should issue valuation guidelines in addition to the ASX and NCSC requirements. My experience in preparing independent expert reports has been that "best practice" sets higher standards than any laws or requirements set out by the ASX and NCSC. This is where I see the AusIMM coming in.

Jan 1, 1989

A system of decline mine entries and ore haulages has been designed for the Mt. Windarra mine. This choice was controlled by the nature of the ore- bodies, the mining method chosen and the capital cost of development re- quired for production. Mine entries are being driven 4 metres high and 5.5 metres wide and decline at 11%, to accommodate diesel haulage vehicles of 35 tonnes capacity. The.Mount Windarra haulage system and the design parameters used in its selection and design are discussed, and general design and selection parameters are proposed.

Jan 1, 1973

Samples weighing 0·5 g are fused with potassium hydrogen sulphate, leached with 4 ml. of 4 M nitric acid and diluted to 10 ml. with water. An aliquot of the clear solution is treated with EDTA, adjusted with ammonia to give a pH value between 5 and 6 treated with cyanide, ammonia, and sodium diethyldithiocarbamate then bismuth diethyldithiocarbamate is extracted with chloroform and the yellow colour is compared visually with a standard series. Alternatively, the bismuth complex may be filtered on a "Millipore" filter and subjected to x-ray fluorescence spectrometry.

Jan 1, 1971

The South Wall Base Case Design for Mt Whaleback indicated that along portion of the South Wall a cigar shaped synclinal keel, contain- ing 14 800 m', of Dales Gorge Member ore, would be unkeyed and bed- ding undercut. The dry "ore lobe" slope would be 60m high, unbenched and approximately 170 m long. Stability analyses using specific design parameters established that artificial reinforcement could stabilise the slope and be economically justifiable. To implement the slope design reinforcement for such a complex geo- logical structure it was necessary that all functions pertaining to slope excavation be integrated and co-ordinated within the Mine Geotechnical Section. The objective being to minimise slope damage and thereby optimise the effect of design reinforcement. The successful excavation of the ore lobe in 1988, three years after the introduction of artificial reinforcement, was achieved by strictly main- taining the following parameters. 1. Accurate Geological Model 2. Controlled Perimeter Blasting 3. Controlled Excavation Techniques 4. Quality control -Reinforcement Installation 5. Additional Reinforcement.

Jan 1, 1989

By Weller K. R, Trahar W. J

The Mount Lyell Company treats some 2.3 million tonnes of ore annually, mined from a number of orebodies on the field. Blended ore is essentially a quartz-sericite schist containing about eight per cent of pyrite and three per cent of chalcopyrite. Four-stage crushing and two-stage ball milling are followe by selective flotation for copper. Pyrite is recovered from the coarse fraction of dopper circuit tailings. A detailed survey involving examination. of 82 samples provided information on the size- by-size behaviour of the principal ore constituents throughout the grinding, copper flotation and pyrite flotation sections of the circuit. Results for the copper and pyrite flotation sections are presented in the paper, and it is shown that over half the copper reporting to the final tailing (5% of copper in the feed) is plus 100 microns and that the major pyrite loss is in the desliming cyclone overflow. These data proVide a basis for assessing potential methods of circuit optimisation.

Jan 1, 1977

By Adams R. W

A commercial scale strong acid leaching plant was commissioned at the Risdon plant of the Electrolytic Zinc Company of Australasia Limited, a member of the Pasminco group of companies, in February 1988. It was designed to upgrade the lead/silver residue product from 26% to 37% lead, thus improving the realised value and increasing overall zinc recovery. This in turn enables the lead smelter treating the lead residue to improve its productivity. The equipment consists of two 250m3 bricklined leach tanks and two 6.5m diameter stainless steel high rate thickeners operating with counter current washing at 95¦C and 120 g/L H2SO4. The present filtration and drying equipment is also being replaced. Two 35m2 squeeze action filter presses equipped with horizontal plates are being installed, together with ancillary drying equipment to ensure that the final product complies with the transportable moisture limit. The new filtration plant is due to be commissioned in mid-1989.

Jan 1, 1989