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By Manca PP, Rossi G, Lipari D, Knissel W, Massacci G, Licheri P, Fioravanti E, Helms W, Lai S

Mining of the massive complex sulphides orebodies of Campo Pisano and San Benedetto (Southern Sardinia, Italy), dates back to before Christ, though their exploitation with modern technology was only resumed in the middle of the last century. The increasingly adverse economic conditions and the weakness of both country rock and mineralization were concurrent in forcing the Company's management to develop suitable high-productivity stoping methods. In the framework of a project fi- nanced in part by the Commission of the European Coranunities, and carried out in cooperation with the Universities of Cagliari (Italy) and Clausthal (Germany), a cemented backfill stoping method was developed. The concrete mix preparation and placement facil- ities, study of the cement mixes and endeavaa-s to optimize the methods are described. The monitoring system for measuring stress evolution within the emplaced backfill is illustrated together with the successful replacement of part of the Portland cement

Jan 1, 1992

By Cunningham DA

Plant based on the New Jersey refluxing process has been installed at the Cockle Creek Works of Sulphide Corporation Pty. Limited to refine portion of the zinc output from the Imperial Smelting Furnace. Refining of impure zinc is effected by distillation in two stages in vertical columns compris- ing a series of superimposed silicon carbide trays designed such that ascending zinc vapours and descending liquid metal are in intimate contact. In the first stage the volatile elements (zinc and cadmium) are distilled from impure metal, condensed, and while still molten, fed to the second column in which zinc is boiled under controlled refluxing conditions to remove the cadmium content. The resultant zinc has a purity in excess of 99,995 per cent zinc. The availability of refined zinc has made possible the production of a number of zinc alloys for marketing in addition to the basis grades of zinc.

Jan 1, 1969

By Savidis GM

This report outlines the experience of Australian mining operators in dealing with large falls of roof in advance of longwall supports. Previous and present techniques used for face stabilisation have been described, together with an analysis of [actors which contributed to the roof falls, in each case. Attention has been focused on the stabilisation technique used at Angus Place Colliery, where the face coal was successfully reinforced by injecting chemical foam into Lite fractured areas. Although considered to be one of the worst roof falls in the history of Australian longwalls (60 m in length and up to 10 in in height), it was subsequently possible to resume cutting in Longwall 9 two weeks after the foam injection. A preliminary schedule of work involving the 'strata replacement' method, is also discussed. It has been based substantially on experience gained with cementitious supports, principally by ACIRL and Monier Resources.

Jan 1, 1986

The major greenhouse gas emissions from coal combustion are leading certain people to argue for uranium mining and nuclear power to meet these challenges. Although there is considerable rhetoric around promoting nuclear power as 'green' due to its relatively lower greenhouse intensity, a careful study of the data is required to properly underst and the sustainability of uranium mining and nuclear power and whether it can or should have a role in future energy supply scenarios. This paper presents detailed research and analysis on uranium mining, focusing on known economic resources, production, solid waste and water management, pollution issues as well as energy and greenhouse aspects. A particular focus will be on Australia and its operating and prospective uranium mines. Overall, the paper raises fundamental issues concerning the long-term sustainability of uranium mining and nuclear power and documents the greenhouse intensity of uranium production and the nuclear fuel chain.

Jun 1, 2010

Blasting is at first sight a simple operation. Blastholes are drilled in a suitable pattern, primed, charged with explosives, stemmed and finally the blastholes are connected together to give the appro- priate initiation sequence. In practice the action of the explosive takes place in the interior of an opaque rock mass and only the external faces of the rock mass to be blasted can be easily observed. The mechanisms of blasting cannot be observed directly and must be inferred from observations of the free face and meas- urements of particle movement made by sensors grouted into holes in the rock. Explosives react very rapidly, in some microse- conds and any instrumentation of a blast must have an appro- priate response time. In many mines several rock types can be encountered during production. Several different methods of blasting may be necessary to cope with changes in rock types espe- cially if weak rocks are found in the final limits. This also makes it difficult to be sure of getting a large enough volume of rock to perform a series of reproducible experiments. While there appears to be agreement on the mechanism of blasting it is unlikely that a detailed understanding of how all the variables that can be encountered in practice affect blasting. The ease of assessment of the results of blasting depends on the criteria used. It is easy to isolate, measure and count over- size boulders. It is however difficult and costly to get a complete fragment size distribution. Methods used to assess blasting can vary significantly from site to site and the ways in which data are collected, presented and analysed will vary making compari- son of blasts at several locations very difficult. The definition of an optimum blast can also vary with the type of blasting. The muck-pile profile and fragmentation will be different for a dragline to that for a truck and shovel operation. In order to be able to assess blasting the mechanism has to be understood. Th rock and explosive properties considered rele- vant to blasting can then be determined. This paper discusses the generally accepted mechanism of blasting. The on-site observa- tions to assess blasting are-then listed. The mechanism of blast- ing can also be modelled and the use of modelling to optimize hlaetina ie thPn Aiernccarl

Jan 1, 1988

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

It is a great pleasure to accept the invitation to open this Joint Conference of The Australasian Institute of Mining and Metallurgy and the American Institute of Mining, Metallurgical and Petroleum Engineers. It is the second invita- tion to participate in the Conference that I have accepted. When it was being planned, the President wrote to me to invite me to be one of the speakers in one of the technical sessions. I wrote accepting the invitation, and when my present appointment was announced, there was a nice question about the earlier commitment. The upshot of it all is revealed by the present arrange-., ment: you get me as opener and the more sub- k stantial speech is to be given by Professor Rupert Myers, my friend and former colleague who will not only be able to contribute from the standpoint of an experienced University admini- strator, but will also speak from a professional training and experience much more closely relat- ed to the concerns of this Conference than I possess. Though it is a joint meeting and the speakers will include distinguished figures from North America, I hope that since the venue is Australia, I may be permitted to speak with an Australian emphasis. I have recently had a brief glimpse of part of the Australian mining industry. In the course of a journey through the Northern Terri- tory we-my wife and I-visited Gove, Groote Eylandt, and Tennant Creek. At Gove we saw the remarkable Nabalco development in bauxite and alumina and the very impressive town plan- ning development at Nhulunby, the major man- ganese enterprise at Groote Eylandt, and the gold mining activity at Tennant Creek. For lay people like us, a gold pour was arranged at Tennant Creek and that was a spectacular sight. On earlier occasions, and before my present appoint- ment, I had visited Mount Isa and the Utah fields, and Queensland Alumina at Gladstone. I have not seen the cradle of the modern mining industry, using cradle perhaps loosely, at Broken Hill, nor have I seen the great Pilbara develop- ment, nor Weipa. This is not for want of generous invitations and it is my hope that within the not too far distant future this will be rectified. Then there is the Australian petroleum industry and the great industries which make their products from the minerals mined which I have yet to See.

Jan 1, 1978

Boliden's Aitik open pit mine located in northern Sweden, will expand yearly production from 18 to 36 Mt of copper ore during 2010. The expansion includes a new concentrator located 3 km off the pit, two new apron feeder crushers connected to the ore storages and concentrator with conveyors at a total length of 7 km. The concentrator has two lines with autogenous grinding mills and flotation each with a capacity of 18 Mt. Consequently the mine operation has to be expanded. To achieve this, there was several parallel processes ongoing; investments in new equipment, reorganisations, improvement projects in maintenance and productivity. The Aitik 72 project was started to coordinate productivity and mine planning activities. Initially the long-term strategic mine plan called for a total movement of ore and waste to increase from 40 to 72 Mt but has today increased to 89 Mt. Productivity projects have been focused on load and haul activities such as fuel economics (Renstr÷m, 2007), effective utilisation of trucks partly by increasing manning levels, improved dispatch and shovel utilisation focusing on optimal use of loader/truck fleets. The eco-driving projects previously reported have reduced fuel consumption significantly. This project is also a productivity project, as part of the fuel savings comes from minimising idle engine hours due to queues, break-downs, operator shift changes, etc but also from under and overloading. Effective use of truck fleets, ie transport of material, loading and dumping, thus excluding non-productive engine hours has been an essential motivator when deciding the size and number of new trucks required. This has resulted in a reduction of the investment with at least one but maybe two trucks. A third major project was to optimise shovels together with truck fleet by improving short-term, up to two weeks, detailed planning of drilling and shovel activities. Daily scheduled breaks for maintenance as well as planning of shovel utilisation, down to shift level are included in the production plans. KPIs and online production reports, based on real-time monitoring of drill rigs, shovels and trucks, have been an essential tool used regularly to engage and motivate mine production personal and information reporting to management. Several projects have already reached their goals. The projects will continue throughout 2010 to ensure that productivity targets are met.

May 1, 2010

By Bogar D. V

Often dewatering of coal tailings is inhibited by the presence of clays, making it difficult to obtain a high concentration, handleable coal tailings. The net result can be a low concentration, chocolate mousse-like tailings which poses significant sedimentation problems in a tailings dam and represents a significant environmental problem. This paper examines the coal tailings from a Hunter Valley colliery where the tailings properties are influenced significantly by the presence of montmorillonite clay. The influence of the calcium ion on the rheology and sedimentation/compression behaviour of coal tailings . containing montmorillonite already dispersed in water was investigated and revealed that very little improvement in the properties is achieved through the coagulation. The preliminary investigations also indicated that the properties of the tailings were dominated by the presence of the montmorillonite clay. Thus investigations were carried out on the clay alone in water. Investigations into the effect of dispersing samples of the clay alone in water and then coagulating with calcium compared with dispersing the clay directly into calcium ion solutions at the same calcium concentration showed huge improvements in the rheology, compressibility, sedimentation rates and sediment solids concentrations for the second dispersion method, which can be called controlled dispersion. The sediment volume fraction solids improved by more than a factor of two by controlling the dispersion of the clay. The reason for the improved properties lies in the suppression of the initial swelling of the clay, which is initially in the sodium form, thus maintaining the parallel orientation of the platelets that is present before wetting. While swelling is being suppressed, ion exchange occurs with the calcium which results in permanent swelling suppression. In the uncontrolled dispersion case where swelling is uninhibited, the clay is allowed to break up and despite the powerful coagulating nature of the calcium ion, the random orientation of the high aspect ratio clay platelets inhibits effective coagulation. Results indicate that the initial electrolyte concentration of the washwater in a coal washery is the critical factor controlling the dewaterability of clay tailings and the handleability of the dewatered product rather than the actual act of coagulating with the calcium ion. The results presented are of clear practical significance and indicate how coal tailings problems where clay dominates the rheology can be improved considerably by knowledge of the surface chemistry.

Jan 1, 1995

By F Owusu-Mensah, F Nsarko

The Obuasi Mine has been in operation for the past 100 years. Within this period a host of mining methods have been employed during its various changing phases. The causes for the changes are multiple, ranging from orebody capability, dynamics, technology, safety, economics, production levels, geotechnical, etc. Mining methods applied have also ranged from timber square setts, conventional cut-and-fill, shrinkage stoping, sublevel caving, mechanised cut-and-fill, longitudinal and transverse open stoping, and lately longitudinal retreat mining method (LRMM).The Obuasi Mine has a world class underground resource potential of about 27 million ounces. A third of this is in reserve. The orebody is subvertical, with multiple lodes and with widths ranging from 1.5 - 40 m. It stretches along 8.5 km N/S strike. A cursory statistical analysis on the various resource ore widths show a higher percentage (60 per cent) of the resource being within the range of about 1.5 - 10 m. Thus a high proportion of the resource is amenable to narrow vein mining methods. This has enhanced the choice of LRMM as an alternative method. Also, with the increasing cost of mining over time, LRMM, which eliminates expensive and time consuming footwall drives, will become far more advantageous as compared to the transverse/longitudinal open stoping methods which currently constitute about 40 per cent of the total underground production.The LRMM currently delivers about 20 per cent of the Obuasi Mine production. Plans are underway to improve total underground production from the current average of 160 000 tpm to about >180 000 tpm in the next few years and about 20 - 40 per cent more in the foreseeable future. Recoverable ounces will also increase from 26 000 to about 34 000 ounces per month. The proportion of LRMM in the business plan is also expected to increase from the current modest level to about 50 per cent. Despite its tremendous advantages in terms of economics and productivity, the method is facing some challenges during its inception. The objective of this paper is to highlight some of the emerging challenges and their effect on production and offer possible remedies that could mitigate the current and possible future challenges.CITATION:Nsarko, F and Owusu-Mensah, F, 2012. The Obuasi mine - longitudinal retreat mining method application and challenges, in Proceedings Narrow Vein Mining 2012 , pp 59-70 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Mar 26, 2012

By G Swayn, J Harpley

Spent Pot Lining (SPL) is an unavoidable waste product of the electrolytic process in the smelting of aluminium. SPL is considered to be a hazardous waste in various countries because it contains significant quantities of absorbed fluoride along with traces of cyanide. The disposal of SPL has been primarily in landfill because of difficulties in the development of a successful techno-economic SPL treatment process. Increasing concern about SPL landfilling practices is resulting in regulations in some countries to ban this form of disposal. As a consequence stock piling of SPL is occurring in an increasing number of countries pending the development of a successful treatment process. In 1992 Portland Aluminium investigated a SPL treatment process based on pyrometallurgical technology and approached a company called Ausmelt in Dandenong, Victoria, Australia, to ascertain if their submerged lance technology furnace process would be suitable to treat SPL. This approach resulted in Portland Aluminium, Alcoa and Ausmelt personnel working together on trials during 1992 - 1994 in Ausmelt's demonstration furnace to produce a slag and off gases containing hydrogen and sodium fluoride. The trials also evaluated the effect of oxygen enrichment to increase the concentration of hydrogen fluoride and the ability to recycle sodium fluoride particulate produced as the furnace off gases cooled. Other separate developmental work by Portland Aluminium, using CSIRO, resulted in the production of aluminium fluoride from the offgases from the pyrometallurgical process. The results of these projects influenced Portland Aluminium and Alcoa to authorise up to $A26 400 000 in December 1995 to construct a Research and Development processing facility to treat SPL at Portland. The Portland Aluminium unique gas treatment process converts hydrogen fluoride in the furnace offgases to aluminium fluoride in a multistage fluidised bed reactor, designed by Portland Aluminium. The aluminium fluoride produced in the reactor has been trialed successfully in Alcoa's aluminium smelting process as a replacement for imported aluminium fluoride. The installation in mid-2001 of a slag granulation system has enabled the production of a granulated vitreous slag having leachability qualities that conform to Victorian EPA criteria for the unrestricted use of the granulated vitreous slag from 'The Alcoa Portland SPL Process'. The unrestricted use of the granulated slag successfully completes 'The Alcoa Portland SPL Process' which treats the hazardous waste material SPL and converts it into non hazardous useful products.

Jan 1, 2002

By H Holmstr÷m, A Widerlund, M Lindvall

At the Kristineberg mine in northern Sweden sulfide-rich, poorly buffered tailings have been deposited in five impoundments located along a valley. The tailings area has been remediated by applying a combination of till cover and flooding. A new method that has been applied to parts of the tailings area is to raise the groundwater table to saturate the tailings. By increasing pH by liming, thereby reducing metal mobility, a water-covered downstream impoundment is functioning as a trap for metals released from the till-covered impoundments upstream. In this paper the efficiency of the system approach used at Kristineberg is evaluated. Progressive changes of water quality from groundwater in the upstream till-covered impoundments via surface streams within the pond system to the water in the flooded downstream pond and its outlet are described. The results show that the system approach used in the design of the remediation programme of the tailings works as planned. The sulfide oxidation rate has decreased in the till-covered impoundments, and, as predicted, there is an out-flush of old weathering products which will continue for at least five years before equilibrium is reached. The groundwater in the till-covered impoundments has high concentrations of Fe, S, Mg and Zn, but the concentrations of metals such as Cu, Cd and Pb are much lower than before the till cover was applied. The limed pond is an efficient metal trap, with concentrations as high as 2.76 per cent Zn and 0.28 per cent Cu in the calcite-gypsum sludge developed on the tailings. The outlet from the flooded pond, and thus from the whole mining area, releases small amounts of metals to the recipient in comparison to the amounts that are in movement within the mining area. During the period May to November, 13 kg dissolved Zn and 111 kg particulate Zn was released to the recipient, whereas a total of 27 000 kg was transported by surface water within the impoundment system during the same period. The results from Kristineberg show that when planning remediation of complex mining areas, it is useful to adopt a system approach in which different methods are used on different parts of the area along the flow path of the drainage water, whereas flooding of the downstream pond is the key to the performance of the entire system.

Jan 1, 2003

The relative contents of Au in a series of groundwaters from water bores and exploration drillholes, located near Au mineralisation at Stawell, Ararat, Clunes and Ballarat in central Victoria, were found to be clear indicators of relative proximity to mineralisation in each region. In the same series, waters with anomalous concentrations of As distinguished Au mineralisation in bedrock from secondary palaeochannel or deep lead Au. In some groundwaters, acid pH values, high sulphate relative to other anions and detectable concentrations of reduced Fe reflected sulphide minerals associated with bedrock Au deposits. Close to some ore zones, locally anomalous groundwater concentrations of Mn reflected the Fe carbonates that accompany Au in these zones. In addition, elevated concentrations of other elements provided geochemical signatures for the aquifer lithologies associated with Au mineralisation. Of the lithological indicators Co and Ni, U and F, Rb and Cs, Mo and Li respectively reflected the mafic, felsic, potassic rich and quartz zones, principally in the Stawell mine area and to a lesser extent those at Clunes and Ballarat. Mineral stability diagrams, based on groundwater activities of K', Mgt and H`, demonstrated that a series of ore-zone groundwaters follow a trend that conforms with a biotite to sericite to kaolinite alteration sequence in aquifer mineralogy. This is a good illustration of how present day groundwaters under surface conditions can reflect the underlying mineral sequences that accompany these Au deposits. These observations around mine areas suggested that groundwater chemistry could provide a rapid and cost effective method of assessing prospectivity for An of extensive areas in central Victoria. To test this, regional sampling and analysis of 355 bore waters was carried out in a region where potential An bearing rocks are covered by Mesozoic and Cainozoic sediments and volcanics. Contoured elemental abundances and distribution from this program indicated five zones in which groundwaters had elevated Au and despite the cover, appeared to reflect variations in the underlying Palaeozoic rocks. Three of the Au zones could be in terrain affected by extensions or intersections of faults, a feature in common with the established Au deposits in central Victoria. Reflections of concealed Palaeozoic rocks included a possible zone of concealed Cambrian greenstones that coincides with three of the Au zones. It was concluded that each of the An zones represents a groundwater Au focus that warrants follow-up exploration.

Jan 1, 1997

The Fosterville Goldfield has a modest past production (0.3 Moz), but is now emerging as a significant goldfield (2.6 Moz resources, including 1.1 Moz reserves) with implications for exploration within the orogenic gold provinces of the Lachlan Fold Belt in Australia and the Buller Terrane on the West Coast of New Zealand. Fosterville is located 20 km east of the 22 Moz Bendigo Goldfield and is geologically similar to other Victorian goldfields, however the nature of the gold mineralisation differs significantly. Unlike other Victorian goldfields there is no primary free gold at Fosterville. The gold occurs in as sub-SEM resolution particles, in fine grained arsenopyrite and pyrite crystals disseminated through the wall rocks in fault controlled quartz-carbonate vein stockworks. The Fosterville Goldfield comprises several subparallel fault systems striking 340¦ over an exposed strike length of 28 km. The Fosterville Fault system comprises the west dipping (~75¦), reverse Fosterville Fault and several less steeply west dipping splay faults formed in the footwall. The splay faults extend through east dipping beds from the Fosterville Fault across a syncline hinge into stratigraphically controlled positions on the west dipping limb of the syncline. The predominant control on gold mineralisation is the intersection of the bedding on the east dipping limb of the syncline and faulting which, is in turn controlled by the orientation of bedding in the west dipping limb of the syncline. The mineralised shoots are thus subparallel to the plunge of the syncline and show a plunge reversal about a culmination in the footwall syncline. This provides a very powerful exploration tool with bedding û cleavage intersection lineations measured in outcrop and oriented core indicating the plunge of mineralisation. Mineralisation is also, albeit less commonly, controlled by local variations in strike and dip of the host fault and at metre scale by host lithology. As the faults and mineralisation are stratigraphically controlled, detailed stratigraphic interpretations are essential to provide control on structural interpretations for resource modelling. Compared to other Victorian goldfields, the Fosterville Goldfield formed later (385 Ma to 360 Ma cf 440 Ma), at lower temperatures and at shallower depths (fluid inclusion studies of Au associated veins suggest vein formation at ~270¦C (range 140¦C to 385¦C) and at depths of 2.6 - 5.7 km).

Jan 1, 2005

By N McIntyre, M Edraki, K R. Jain

From the very early stages of a mining project, there is a requirement to predict the chemistry of potential drainages from various waste storage facilities. However, using relative masses to scale chemical loads measured by laboratory kinetic tests, such as humidity cells, to a full-scale mine waste facility will lead to concentration predictions that are unrealistically high for many dissolved constituents. Scaling factors developed by various procedures indicate that laboratory rates may range anywhere from two to eight times (Drever and Clow, 1995) to 100 to 1000 times faster than field rates (Malmstrom et al, 2000; Smith and Beckie, 2003; Ritchie, 1994). This contribution will use previous works and secondary data to discuss the significance of scale in water quality predictions. It will also present primary experimental geochemical and numerical data from a column leaching study on coal spoils to discuss potential applications of mesoscale leaching tests in bridging the gap between bench-top tests and field scale trials.

Aug 2, 2023

TasmaniaÆs mineral endowment has resulted in 600 abandoned mines. Commonly occurring sulfide minerals such as pyrite and pyrrhotite can generate acid drainage and also mobilise heavy metals. Due to the StateÆs steep topography and high rainfall the risk to water quality in the receiving environments is significant. Surface waters impacted by abandoned mine sites typically showed high sulfate and metal distribution at pH range of 2.0 to 5.0. The impacted waters generally showed metal distribution in the high acid/extreme metal to low acid/low metal range. A similar trend was shown by sulfate distribution. Elevated levels of metals and sulfate downstream from several abandoned mine sites indicated that active sulfide oxidation is occurring and these sites may be potential point sources of metal pollution in the receiving environment. High correlation between metal and sulfate distribution suggests that sulfate distribution in surface waters may be used as a cost-effective investigation tool for preliminary assessment of acid drainage.

Jan 1, 2003

This paper will present a series of examples drawn for a number of Australian and international mining companies showing their recent attempts at optimisation of capital. The results will be compared, contrasted and discussed in the light of six factors that are considered crucial to any analysis of this subject. The factors are: ò the market availability of capital ò the corporate strategy and performance ò the capital structure appropriate to the corporate positioning within the industry ò the parties to the optimisation equation ò the measurement of success of optimisation The factors above are not mutually exclusive and are generally interdependent so their order above is of little consequence. Other papers presented at this conference will dwell in more depth on all these factors, therefore this paper shall focus only on the key elements of each factor. By way of summary this paper will also address the issue of the longevity of optimisation - is it temporary or long term.

Jan 1, 1991

This case study reviews the Charters Towers vein systems for potential open-pittable bulk tonnage resources in' stockworks, horsetails or multiple vein systems. The review summarises district evaluation undertaken over the last 18. months before specific drill evaluations.Historical mining from Charters Towers includes in excess of six million ounces at head grades of one ounce per ton. A review and structural studies show most veins as simple, tight systems without significant horsetails or multiple sheeted systems. In addition the normal gold content of the pyritic selvages ('formation') is low and disseminated gold is virtually absent. Shoots are structurally controlled on vein/vein and vein/contact intersections. Historically high production vein systems are hosted by .,.t o of the younger granodiorite phases.Geometrically most veins dip shallowly (from 45¦ to as low as 200) in a north or easterly quadrant. Flat dips permit open cut exposure of considerable vein area at a favourable stripping ratio.

Jan 1, 1986

The concept of Extreme Gravity is the latest in a series of innovations that has seen a re-introduction of gravity recovery systems into the milling operations of most gold mines. Traditionally, most gold milling circuits are designed around flotation and cyanidation requirements, with the gravity circuit being fit in where possible. Extreme Gravity takes the novel approach of optimising the circuit, in order to maximise recovery by gravity. In some cases gravity systems can achieve high enough recoveries to eliminate the need for chemical systems such as cyanidation or flotation. There are many advantages of such a system including low capital cost, reduced permitting, improved cash flows and much reduced environmental issues. There are a number of plants currently in operation which utilise this new methodology. The Muzhievo project has been an example of a recent turn key success achieving recoveries from 85 - 90 per cent. The details of this project are examined as a case study.

Jan 1, 2000

The occurrence is recorded of an indurated kaolinite claystone associated with red-beds in the Burralow Formation of the Narrabeen Group near Putty toward the north-west of the Sydney Basin. The claystone is similar in composition and structure to the Garie Member which immediately overlies the kaolinite-rich "chocolate shales" of the Bald Hill Claystone in the sofrthern part of the basin. Unlike the Garie Member however, the Putty claystone is lenticular and the associated red-beds contain abundant quartz and illite in addition to kaolinite.The source of the Putty claystone is not known. Possibly it is an extension of the Garie Member to the northwest or alternatively, it may represent fine outwash of detritus that produced the coarse-grained kaolinite claystones in the Triassic succession of the Merrygoen-Binnaway area. Although the material is suitable for the manufacture of aluminous refractories the limited size of the deposit precludes its commercial exploitation.

Jan 1, 1971