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By Ellyett C. D

This paper is an appraisal of the application of optical-mechanical-scanners operating in the non-photographic infrared wavelengths (3-5 and 8-14 micrometres) to mineral exploration in Australia. Pundamental principles of the operation of airborne scanners and their data output, thermal infrared (IR) imagery, are summarized. It is contended that thermal IR imagery often depicts the distribution of electromagnetic radiation emitted from earth surface materials better than aerial photographic techniques because optical-mechanical-scanners sample wavebands which straddle the wavelength peak of earth surface emission. To illustrate the geological features which such scanners can depict, reference is made to thermal IR imagery of a number of areas in Australia where the technique has been applied for exploration purposes. The images demonstrate that this technique can detect structural features like faults and folds, stratigraphic data by distinguishing different rock types and delineating their

Jan 1, 1973

By Ryan CG, Both RA, Mernagh TP, Large RR, Hunns SR, Gemmell JB

The Hellyer and Mt. Chalmers deposits are mound-style volcanic-hosted massive sulphide (VHMS) deposits in Australia. Textural, petrographic and microthermometric investigations of fluid inclusions in the Hellyer stringer system indicate that Type I, primary, liquid-vapour inclusions are 10-15 pm in size, and yield homogenisation temperatures of 170-220¦C in early 2A veins, 165-322¦C in main-stage 2B veins and 190-256¦C in late-stage 2C veins. These data suggest a waxing and waning thermal history. However, the average salinity remained between 8-11 NaCl equiv. wt % in all Stage 2 veins. At Mt. Chalmers, Type I inclusions up to 20 ¦m are found in quartz from the mineralised zone, and these inclusions yielded homogenisation temperatures of 160-268¦C and salinities of 5-8 NaCl equiv. wt %. Laser Raman spectroscopic (LRS) analysis indicates the presence of CO2 (

Jan 1, 1995

By Grieve BL

For a blast to be successful both heave and fragmentation must be satisfactory. Amongst the major factors influencing blasting are rock properties. This paper considers the effect of rock density on fragmentation and heave and how better results can be achieved in dense ore. Patterns which increase the heave and give better fragmentation are examined and the application of theory to the blasting of the Dales Gorge Haematite, density 4.2 - 4.5 t.m.,3 is then discussed. The drilling of staggered patterns instead of square patterns and a shallower 'V' initiation system has given better heave and fragmentation. This practice has re- duced backbreak and improved toe conditions; these factors have resulted in increased load ing rates. There has been an increase in the lump/fine ratio which is consistent with the theory although other factors such as ground conditions could have contributed.

Jan 1, 1978

By Richards AB

This paper describes the ENVIB software system, which has been developed in a menu-driven form to aid in the monitoring and assess- ment of vibration caused by routine blasting operations and for research and development projects. The paper gives an outline of the capabilities of the system by means of screen-dumps and a general descriptive commentary. Technical notes have been added where appropriate to further describe the system. The ENVIB system covers: I . Analysis of blast vibration results, and preparation of graphs and tables for control and estimation purposes. 2. Effects of atmospheric conditions and air vibration propagation. 3. Real-time acquisition and analysis of analogue blast vibration sig- nals directly from transducer equipment or from FM tapes. 4. Analysis of digital files on floppy disc or transferred via the serial port from blast monitoring equipment. 5. Transfer function calculation for structure response studies.

Jan 1, 1988

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

By Woodhouse M, Drew GJ, Andrews DL

The Honeymoon nickel sulphide deposits are located 45 km south of Wiluna in the 2.7 Ga old Agnew-Wiluna greenstone belt in the NE of the Yilgarn Craton (Figure 1). Both disseminated and massive nickel sulphide deposits are present within the deformed and metamorphosed Honeymoon Well ultramafic complex and are hosted by distinctly different komatiitic lithologies (Gole et al, 1996). The deposit styles and host rock lithologies reflect formation in markedly different volcanic settings. Metasomatism has affected parts of the ultramafic complex, including sections of the sulphide deposits, and has had a marked impact on mineralogy and bulk sulphide compositions. The present indicated resource is 158 Mt at 0.71 per cent Ni in four deposits (0.4 per cent Ni cut-off, resources above 300 m depth, December 1995). This includes a massive sulphide resource of 2.5 Mt at 3.36 per cent Ni.

Jan 1, 1996

By C J. McKenzie, D F. S Natusch

The lignite deposits of Southland and Central Otago are a potentially significant energy resource for New Zealand. Current estimates put the technically mineable lignite reserves at 5.3 billion tonnes, which is more than ten times the energy content of the Maui gas field. Thus, the strategic importance of the lignite as a basis for maintaining current levels of self-sufficiency in transport fuels, or as a possible feedstock for electricity generation, deserves full investigation. The full extent of the lignite deposits was only recently established as the result of a scout drilling programme undertaken between 1976 and 1979 by the Mines Division of the Ministry of Energy, and from companion studies conducted by the New Zealand Geological Survey. These discoveries of large quantities of lignite coupled, at the time, with rising oil prices, raised the prospect of their development for production of transport fuels.

Jan 1, 1983

In the first half of 1987 two four metre diameter raise bore holes were drilled at the Olympic Dam Project for ventilation purposes. The first hole, 476 m x 4.0 m diameter was completed in 96 days, 35 days longer than scheduled. The second hole, 407 m x 4.0 m diameter was completed in 39 days, 2 days ahead of schedule. Both holes were drilled with a Robbins 121R rig using a Wirth head that consisted of 23 face cutters and 5 side cutters, supplied by a contractor. After the sites were located a grouting program was carried out to help seal off an aquifer between 160 and 180 m below the surface. Cavities discovered in the Andamooka Limestone were filled with concrete. Because of the overlying dune sands and clay both sites were pre-sunk to the competent Andamooka Limestone.

Jan 1, 1988

The patterns of occupational health pro- grammes in mining and metallurgy are discussed in relation to the impact of chemical and phy- sical hazards on the health of workers in these fields. Mining and metallurgy are quite unique in that their history and practice embrace virtually the total spectrum of occupational health problems and hazards as understood today. A panoramic concept of the possibilities would comprise the following fields: 1. The Pneumoconioses or Dust Diseases. 2. Physical stress due to exertion, and exposure to heat and humidity in deep mines. 3. The toxic effects of various metals in the mining of ore. The most obvious are lead, mercury, arsenic and manganese. 4. Diseases associated with mining. These include leptospirosis, ankylostomiasis and mycotic infections of the skin. 5. The Beat Diseases of coal miners working in narrow seams. 6. Illumination problems and miners nystagmus. 7. Geographical and climatic problems

Jan 1, 1975

The patterns of occupational health pro- grammes in mining and metallurgy are discussed in relation to the impact of chemical and phy- sical hazards on the health of workers in these fields. Mining and metallurgy are quite unique in that their history and practice embrace virtually the total spectrum of occupational health problems and hazards as understood today. A panoramic concept of the possibilities would comprise the following fields: 1. The Pneumoconioses or Dust Diseases. 2. Physical stress due to exertion, and exposure to heat and humidity in deep mines. 3. The toxic effects of various metals in the mining of ore. The most obvious are lead, mercury, arsenic and manganese. 4. Diseases associated with mining. These include leptospirosis, ankylostomiasis and mycotic infections of the skin. 5. The Beat Diseases of coal miners working in narrow seams. 6. Illumination problems and miners nystagmus. 7. Geographical and climatic problems

Jan 1, 1975

By Abdurazakov A, Avrakhov A, Solozhenkin P

The flotation properties of acetic acids modified by xanthic and dithiocarbamic fragment in regards to copper and ferrous minerals and their noble metal containing ores have been investigated. The effectiveness of the co-using of these reagents with xanthates and dithiocarbamates has been shown, ie using of proposed reagent regimes for the flotation of Au-containing ores increases gold recovery by 4.2 - 4.7 per cent, and in the case of flotation of copper Re-containing ores copper recovery increases by 3.3 per cent, rhenium recovery - by 15.5 per cent. The mechanism of the action of these reagents has been explained by the data of UV and ESR-investigations of their sorption on mineral surface. The information on structure and properties of co-ordination compounds of some metals with (Diethyldithiocarbamato) acetic acid were used also for the explanation of flotation mechanism.

Jan 1, 1993

Record quantities of slag were produced in 1974. Slag usage in this year amounted to 8.5 mega tonnes for granulated blast furnace slag and 7.7 mega tonnes for crushed blast furnace and L.D. slags. Granulated blast furnace slag is mainly used as a hydraulic bonding material in the manufacture of cement and in road construction. These applications give a fuel saving and reduce processing costs. Crushed blast furnace and L.D. slags find application as a substitute for natural aggregates in road construction and in the production of concrete. There is no doubt of ttie interest to utilise to the maximum all the slag produced by the steelmaking industry, thus preventing the build up of slag heaps detrimental to the environment and favouring the disposal of the heaps formed in the past. World wide efforts are being made to achieve this objective.

Jan 1, 1979

The non-metallic minerals, excluding the construc- tions and fuel materials have for long been the "poor relations" of the Mining Industry of New South Wales. The State is self-sufficient in only a few of the industrial minerals and is almost completely lacking in any of the essential fertilizers. The deficiencies must be met by imports from overseas and interstate and in fact New South Wales imports some quantity of every one of the non-metallics. For the year 1967 the value of non-metallic mineral production amounted to only $9,390,000 or less than 3.5% of the States total of $275,900,000 for all classes of mining and quarrying. In roughly the same period, the 1967-68 financial year, New South Wales imported non- metallic materials, valued at well in excess of $22,000,000 from overseas and over $2,000,000 from interstate sources.

Jan 1, 1969

By P Joyce, D R. Cooke

The Dinkidi Cu-Au porphyry deposit is situated in the Caraballo Range of North Luzon, Philippines. Although comparatively small, Dinkidi has unusually high gold grades with drilling by Climax-Arimco Mining Corporation delineating a 110 Mt orebody at 1.2 g/t gold and 0.5 per cent copper. The magmatic-hydrothermal evolution of Dinkidi records a distinctive switch in mineralisation from an early orthoclase-dominated Au-rich assemblage, typical of alkaline porphyry systems, to a later quartz stockwork Au-Cu rich assemblage typical of most calc-alkaline to shoshonitic porphyry systems. The early orthoclase-dominated stage is associated with the intrusion and crystallisation of a variably-textured clinopyroxene-syenite dyke that contains high-grade (5 - 20 g/t) gold mineralisation. We consider this dyke to be a Cu-Au porphyry analogue to the pegmatites found in granitic systems. Dinkidi is hosted within the Early Miocene Didipio Igneous Complex. This alkaline igneous complex is comprised of early unmineralised gabbros (pyroxene-plagioclase cumulates), clinopyroxenites (pyroxene cummulates), diorites and monzodiorites that have been intruded by a large, weakly mineralised monzonite pluton. The Dinkidi deposit is located at the southern margin of the monzonite pluton. Mineralisation is hosted within a composite monzonite stock. Within the stock, an 800 metres by ~200 metres (plan section) elongate monzonite stock (Tunja Monzonite) predates and hosts the bulk of the Au-Cu mineralisation. This has been intruded by the thin (10 to 30 m wide), highly mineralised variably-textured magnetite- and clinopyroxene- phyric syenite pegmatite (the Balut Dyke). The youngest phases of the composite stock are leucocratic quartz-monzonite porphyry (Quan Porphyry) that grades texturally, in its core, into a post-mineralisation crystal-crowded quartz-syenite (Bufu Felsite). The late-stage Bufu Felsite is inferred to be the source of much of the Cu-Au mineralisation. The Didipio Igneous Complex has intruded a sequence of Late Oligocene alkaline andesitic lavas and volcaniclastics and overlying potassic dacitic lavas. Whole rock geochemical analyses indicate that the intrusions and the potassic dacite lavas in the surrounding region are co-magmatic. Five main stages of hydrothermal activity are recorded at Dinkidi. An unmineralised K-silicate (biotite-magnetite) alteration assemblage formed during the intrusion of the Tunja Monzonite stock. The diorites surrounding the stock display intense selectively pervasive biotite and magnetite alteration and are cut by an extensive network of orthoclase veins and monzonite dykes. The orthoclase-dominated assemblages of Stage Two carry substantial high-grade Au mineralisation and are associated with the intrusion of the Balut Dyke. The emplacement of this unusual dyke is associated with emplacement of pegmatitic clinopyroxene-orthoclase-magnetite¦sulphide dykes, localised orthoclase flooding along the margins of the dyke, and subsequent emplacement of an extensive orthoclase-actinolite-bornite¦chalcopyrite-fluorite stock-work, and development of widespread selectively pervasive actinolite alteration. The main stage of Au-Cu mineralisation at Dinkidi is intimately associated with the intrusion and initial crystallisation of the leuco-monzonite porphyry (Quan Porphyry) and emplacement of a quartz stockwork followed by carbonate-sulphide deposition and intense vein-related sericite and carbonate alteration assemblages. The quartz stockwork formed synchronous with the deposition of the æBugoy QuartzoliteÆ, a very coarse grained mosaic of quartz-orthoclase-perthite along the walls and top of the crystallising intrusion. Stage 3 was terminated by crystallisation of the distinctively vuggy Bufu quartz-syenite phase in the core of the intrusive, with associated antiperthite veins, and the subsequent explosive brecciation of the roof of the intrusion, forming a quartz-fragment breccia from the Bugoy Quartzolite. Retrograde metasomatism during Stage Four was intimately associated with the emplacement of extensive, late-stage carbonate, silica and zeolite veins and widespread selectively pervasive carbonate alteration. The last hydrothermal stage at Dinkidi, Stage 5, is comprised of strong selectively pervasive clay alteration and is associated with fault-related brecciation of the Dinkidi stock.

Jan 1, 1999

By R Schiller, E Chait

"Quebrada Blanca is a porphyry copper deposit located in the Eocene-Oligocene belt of Chilean Andes. Operation started in 1994 and original reserves for supergene mineralisation gave a life-of-mine (LOM) of 20 years; however, further exploration extended the LOM until 2020.In order to optimise the remaining five years of supergene resource, geometallurgical variables were added to the geological model including copper recovery and acid consumption. LOM plan encompasses finishing the phase number eight and the future development of phases 11 and 13.LOM lithologies include diorites, quartz monzonites, porphyries, igneous and tourmaline breccias. Main alterations are: potassic, phyllic and argillic. Litho-alterations or geometallurgical units were formed by combining five types of rocks with three types of alteration. According to statistical information, the deposit can be represented with 12 litho-alteration units, which cover 98 per cent of the LOM volume.For copper recovery information, there were 40 mini-columns tests made in order to obtain kinetics curves and copper recovery values after comparing the difference between head and tails grades. Acid consumption was obtained using an automatic neutraliser device-based test.Historical data said that copper recovery in pads after 450 days of irrigation is around 75 per cent of recoverable copper while acid consumption is around 5.5 kg/t.According to this study, lower recoveries are related with potassic-altered diorites and argillic-altered quartz monzonites with values around 60 per cent. On the other hand, best recoveries could be found in potassic-altered quartz monzonites and argillic-altered diorites with values closer to 80 per cent.It was also revealed that the most reactive rock is the potassic-altered diorite with 12 kg/t of acid consumption. This rock covers approximately 40 per cent from phases 11 and 13.The final geomet model was calculated using a program that relates each geometallurgical unit with a specific copper recovery and acid consumption value. Results found after comparing model-obtained information with historical leached pads data were used to validate the model.CITATION:Chait, E and Schiller, R, 2016. Adding Copper Recovery and Acid Consumption Variables to the Geological Model of Quebrada Blanca, in Proceedings The Third AusIMM International Geometallurgy Conference (GeoMet) 2016, pp 257–266 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Jun 15, 2016

By R C. Cox

The Brukunga pyrite mine was operated from 1955 to 1972 to supply feedstock for sulfuric acid production in the South Australian fertilizer industry. Iron sulfide ore, mined by Nairne Pyrite Ltd, was finely crushed and concentrated onsite with the concentrates being sent by road and rail to Port Adelaide for roasting. Waste rock was dumped at two locations on the western side of Dawesley Creek and the tailings pumped to a valley-fill storage facility on the eastern side. Production of 1.5 Mt concentrate resulted in 3.5 Mt tailings and 8 Mt waste rock. The site has been administered by the South Australian Government since August 1977 and a lime treatment plant has been operated on the site since September 1980. Environmental issues arising from these operations included: diversion of Dawesley Creek to accommodate the waste rock dumps; dumping of low-grade sulfidic ore in waste rock dumps at the angle of repose immediately adjacent to Dawesley Creek; exposed fresh sulfide mineralisation in the quarry floor; tailings storage facility dam wall constructed by upstream up-lift using tailings and waste rock; an acid water pond sited on the tailings; and pollution of natural drainage. As a result, the minesite rapidly became a source of acid drainage and potential contaminants (sulfate, Al, Fe, Mn, Cd, Zn, and Ni). Various Government bodies have been responsible for considerable effort and expenditure in reducing the flow of acid seepage and contaminants into Dawesley Creek. Presently an estimated 60 per cent of seepage from the waste rock dumps, tailings storage facility and quarry is collected and neutralised prior to discharge into the creek. Tailings have been covered with waste rock, biosolids, neutralisation sludge and soil and planted with grasses, shrubs and trees to reduce rainfall infiltration. Monitoring of water flows, water quality and riparian ecosystems is on-going. Consultants have been engaged to develop cost-effective remediation techniques and a five-member Brukunga Mine Site Remediation Board was appointed to oversee remediation of the site and liaise with the community. Construction has commenced on the diversion of Dawesley Creek, and plans are being developed for the relocation of waste rock back to the quarry benches and upgrading of the neutralisation plant. The paper examines operations, environmental monitoring, remediation, research and community involvement with respect to current best practice.

Jan 1, 2003

By D Johansson, A Nordqvist, D Mihaylov, N Petropoulos, M Wimmer

The flowability of the caved rock masses in the sublevel caving (SLC) mining method is of great importance. During the blasting process of a SLC-ring, the caved rock mass in front of the ring is compacted. As a result of this compaction, the mechanical properties of the fractured rock masses change, for example void ratio, angle of internal friction and structure of the compacted material. The changed mechanical properties of caved rock mass and recent blasted material can influence the mobility of the material, which could result in blockage of the flow path of the material. The identification of the involved mechanisms in this process can be done after a series of tests in different scales. However, there is no available equipment for dynamic measurements during blasting, ie velocity and displacement of the burden and in the caved masses. This study focuses on the development and laboratory evaluation of a measuring system that can be installed in a borehole. The first trial was carried out in a pillar, where two blastholes had been drilled (9.5 m) parallel with the drift and two measuring holes (16.5 m) were perpendicular to the drift. Two measuring systems were installed in the measuring holes. Some preliminary tests were carried out to find suitable initiation procedures for the emulsion explosive. Applicability and robustness of the developed measuring system were also evaluated. In addition to these tests, an expandable cement mixture was developed to firmly anchor the measuring systems in the boreholes. The measuring system consists of an anchor, armoured cable in a pulling tube and a protective case. The anchor is equipped with uniaxial and triaxial accelerometers. A potentiometric measuring system for displacement up to 2 m is installed at the end of the pulling tube and connected to an extendable spiral cable. All sensors are connected to a data acquisition system (DAS). Both measuring systems have their own DAS to acquire data in case of failure of one of the systems. The measuring system was tested in laboratory conditions with impact velocity up to 10 m/s. It was proven to be functional and gave reliable results.CITATION:Petropoulos, N, Mihaylov, D, Johansson, D, Wimmer, M and Nordqvist, A, 2015. Design of equipment for dynamic burden measurements, in Proceedings 11th International Symposium on Rock Fragmentation by Blasting, pp 493–500 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 24, 2015

By M Chaudhari, G Manekar, D Shome

MOIL Limited is operating 7 underground and 3 opencast mines in central India and producing annually more than 1.2 million tonnes of various grades of manganese ore. The underground mines are operating at shallow - 90 m at Munsar mine to moderate - 353 m at Balaghat mine below the surface with horizontal cut and fill (HCF) or its variant method of stoping with post filling by hydraulic sand stowing. The Munsar mine is being worked underground since 1903. The opencast mining is presently stopped. The ore body below the opencast quarry is geological continuity of the area excavated in past by opencast method of mining. The underground mining is done through the adit in three levels i.e. 270’L, 220’L, 170’L and by incline in western part of the property at 70’L. The horizontal drift development has been developed in the manganese ore body with sill drive of 5 m above the sill pillar in lower level at 70’L and barrier pillar of 5 m thickness is left for protection of upper level at 170’L. Valuable mineral has been locked in sill pillar and therefore rock mechanics investigations have been carried out. On the basis of study now the drift development has been carried out in footwall rock at 70’L, (-) 30’L and below levels and it has improved the mineable manganese ore by about 20%. In this modified HCF stope design, an alternative fill material in place of sand, which has been developed in house by use of overburden (OB) material after treatment, has been used for hydraulic stowing. The experimental trials find out that the alternative fill material of OB is more compact and forming a non expansion floor for the men and machines in the stope. This will certainly help in the introduction of mobile underground mining equipment for drilling in the stopes and mechanical mucking, transportation and loading of the ROM. The paper presents rock mechanics investigations and future use of alternative fill material of OB for better safety and productivity at Munsar Mine of MOIL. CITATION:Manekar, G, Shome, D and Chaudhari, M, 2018. Modified ground support with alternative fill material for ground control at Munsar underground manganese mine of MOIL Limited, India, in Proceedings The Fourth Australasian Ground Control in Mining Conference (AusRock), pp 58–65 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 30, 2018

By M Lee, J F. Carswell, G Reitsema

"The increased costs of underground mining have continually challenged engineers to identify opportunities for alternatives to common approaches to underground development. Considerations to the design of horizontal and vertical development at the Dugald River mine targeted inclined raise boring to reduce horizontal development, not only to reduce cost, but to allow development activities to focus on other areas of the mine.The mine plan consists of 50 raise bored ventilation shafts, of which 17 have been mined to date with diameters ranging from 3.1 m to 5 m and lengths from 30 m to 200 m. The expected performance and stability of these shafts has been determined using methodologies proposed by McCracken and Stacey (1989) and further refined by Peck, Coombes and Lee (2011). Actual raise performances have been collated and compared to a database of raise bore performances across Australia as presented by Coombes, Lee and Peck (2011), and form a database of raise bore performance specific to Dugald River.Opportunities to further optimise the mine plan and reduce costs have seen the introduction of large diameter-inclined raise bored ventilation shafts to minimise the required lateral development to connect the ventilation circuits. Large diameter inclined shafts ranging between 3.1 m and 4.5 m in diameter have been successfully mined with dips down to 70°, reducing the cost of lateral development by more than $1 million over the next two years.Both technical and operational challenges were overcome in the mining of large diameter-inclined raise bores, including modification of: existing methodologies of assessing shaft wall stability for inclined raises. These included variations of the McCracken and Stacey method and the utilisation of the modified stability number (N’) previously developed for stope performance (Potvin, 1988). both the collar and toe development profiles, including the general set-up of the raise bore drill rig to reduce the operational risks associated with angled raises.Due to their success, inclined raises have been designed for subsequent stages of the life-of-mine ventilation system at Dugald River. This has resulted in significant capital cost savings for the mine.CITATION:Carswell, J F, Reitsema, G and Lee, M, 2017. Vertical and inclined raise bore performance at Dugald River mine, in Proceedings 13th AusIMM Underground Operators’ Conference 2017, pp 313–320 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Oct 16, 2017

By Lawson F

Narrowly sized samples of synthetic chalcocite (Cu2S) and covellite (CuS) were leached in oxygenated acidic sulphate-chloride solutions. Chalcocite leaches via a two-stage process, the first stage being much more rapid than the second. The apparent activation energies are 42.7 kJ/mol and 68.2 kJ/mol respectively, indicating that both stages are controlled by surface chemical reactions. The leaching of covellite is also chemically rate limited with an apparent activation energy of 77.0 kJ/mol. The shrinking core model is able to be used to describe the leaching processes of both copper sulphides. Over 97% of copper present in chalcocite dissolved in 0.5 M H2SO4i0.5 M NaCl solution under one atmosphere pressure of oxygen at 85¦C in three hours from particles 36 pm in diameter while only 85% of the contained copper in covellite particles 13 pm in diameter dissolved in the same solution under the same pressure at 90¦C in six hours. About 99% of the sulphide sulphur in chalcocite was oxidised to elemental sulphur while only 90% of sulphide sulphur was converted to elemental sulphur during covellite leaching. The mechanism proposed for the two-stage leaching of chalcocite particles is as follows: in the first stage, cuprous ions rapidly diffuse to the external reaction surface where they lose electrons and go into solution as a complex cupric chloride ion CuCl'. The second leaching stage begins when covellite is formed at the external surface. Copper again goes into solution as the CuCl* ion. The mechanism for leaching synthetic covellite is similar to the second leaching stage for chalcocite except that the leaching rate of synthetic covellite is substantially lower than that of the newly formed 'covellite'. Elemental sulphur formed on the particle surface retards the dissolution processes during both chalcocite and covellite leaching. The presence of chloride ions is necessary if essentially all of the copper from chalcocite is to go into solution. When no chloride ions were present, the leaching rate of both newly formed and synthetic covellite was almost zero. The main role of chloride ions in both chalcocite and covellite leaching appears to be to disrupt the passivation film of sulphur on the particle surface.

Jan 1, 1989