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By M Hedley

Tracking plant and personnel in mining operations provides multiple potential benefits for a mine including improved safety, better monitoring of production, traffic control and, in the future, monitoring and control of autonomous vehicles. The main techniques currently proposed and used for wireless tracking are radio frequency identification (RFID), signal strength based techniques and time of arrival (TOA) based techniques. RFID provide poor accuracy in the complex radio propagation environment present in mines and does not provide continuous tracking and signal strength techniques. TOA based techniques are potentially the most accurate, however, most current equipment has not realised the potential of the technique. CSIRO has been developing the Wireless Ad hoc System for Positioning (WASP) system for wireless tracking in a range of challenging environments including underground mines and surface operations, where GPS reception is compromised by depth. The system utilises a number of ‘anchor nodes’ that are placed at fixed and known locations within the mine and tags attached to the plant or personnel to be tracked. The system provides an ad hoc data communication network, thus, wired data connections are not required to the anchor nodes. WASP uses transmissions in the 5.8 GHz ISM (instrumentation, scientific and medical) frequency spectrum to perform highly accurate ranging between nodes, hence tracking of tags. The localisation performance of the system depends upon the radio propagation environment, ranging down to 0.1 m in low multipath conditions. In this paper we describe the WASP system and provide results of testing in underground and surface operations showing the system achieving high ranging, hence localisation, accuracy.

Nov 22, 2011

Deep hydrothermal fluids adjacent to the Alpine Fault move through the brittle-ductile transition to mix with convecting fluids from near surface environments. These tectonically driven fluids evolve mainly from prograde metamorphism at mid crustal depth (ten kilometres) within the ductile regime which has been elevated due to advection and drag on the fault plane. Through the onset of brittle failure and cataclasis within Alpine Fault mylonites, fluid overpressuring results in hydrofracturing and localized implosion. Along the length of the Alpine Fault, hydrothermal alteration varies on most levels (Table 1). There is an observable increase in Fe-Mg carbonate between the Waitaha and Taramakau Rivers (between Ross and Aurthur's Pass, respectively). Field studies concentrate to the north east of the Arahura River, noted for historical mineral wealth. Late meteoric fluids move through cataclasite, brecciated mylonite, and other open fractures. Chloritisation is the major retrogressive phase. Dolomitic hydrofracture veins (Figure 1) within coherent mylonites contain iron and copper sulphides. Notably, the enrichment in Fe-Mg carbonate towards the Alpine-Hope Fault intersection seems to represent a region of enhanced permeability possibly related to shallow seismicity, increased regional faulting, footwall variation, strike-slip dip-slip distribution, and/or topographical anomalies (ie saddles). Isotope studies identifying an 'Inboard Fault Gouge Calcite Zone' correlate to values from the Main Divide region at the headwaters of the Arahura River (Browning Pass-Wilberforce Valley) (Becker et aI, 2000). These correlations are supported by rare arsenic anomalies (avg l06 ppm n = 5, Figure 2) found in both regions which suggest that fluids may be interconnected at depth.

Jan 1, 2001

By C J. Corbett

Many underground mining methods rely on blasted rock moving under gravity to report to extraction points. Mining of orebodies dipping between 20 - 45 degrees poses a unique challenge for mining engineers in that a dip angle in this range will not allow the movement of broken ore down the footwall to the mucking horizon utilising gravity. This is more so in cases where the natural angle of repose of the rock is less than the dip of the orebody. Hence, innovative mining systems are necessary to mine this type of orebodies. In Australia, the method of uphole retreat panel stoping is commonly used to extract orebodies in the dip range of 20 - 45 degrees and thickness range of 5 - 15 m. In the Zambian Copperbelt a method called continuous retreat updip stoping is used for orebodies with similar dips and thicknesses. The mechanised form of this mining method has potential for application in the metalliferous mining industry in Australia. In this study a comparison is made between mechanised updip (MUDS) and transverse retreat stoping (TRS) methods. A hypothetical nickel sulphide orebody located in Western Australia is used as case study. The net present value for mechanised updip stoping is seven per cent higher than that of transverse retreat stoping based on the assumptions applied. Initial results of this study indicate that mechanised updip stoping can be adopted for Australian mining conditions. However, lack of significant local experience with the method is often cited as a reason for not adopting it even in situations where it is clearly the best option.

Jan 1, 2003

Around 35 Mt of industrial clays were produced in 1994 under the broad classifications kaolin (25 Mt), bentonite (8 Mt) and attapulgite/sepiolite (2 Mt). A brief review of world sources and markets is given as background to Australia's small production (0.5 Mt) but developing interest in industrial clays. Clay markets have changed significantly over the past 30 years resulting initially in consolidation and expanded production from traditional high-grade clay sources but also creating opportunities for development of new discoveries. Decline in the use of clay-based refractories has come at a time of strong growth in demand for refined kaolin for the paper industry. Traditional markets for bentonite as oil drilling muds, iron ore pelletising and foundry bond sands are mature and in overall decline. Today, bentonite has a significant share of the growth market in pet litter which a few years ago was dominated by attapulgite/sepiolite clays. Both bentonite and attapulgite/sepiolite have good prospects for increased demand in environmental markets both in containment of hazardous wastes and sorption of toxic substances. In common with all industrial minerals, markets for clay minerals will demand continuous improvement in product quality and consistency. Development of new and improved products will be increasingly the responsibility of the raw material supplier. In order to remain competitive, this will require a substantial commitment to quality assurance and research and development. Increase in market share will come for those producers able to translate developments in their products to cost savings or increased competitiveness for the industries that they Supply.

Jan 1, 1997

By M F. Navarrete

During 2003, different plastic stemming plugs (Cats«, Stemtite«, and Taponex«) were tested in EscondidaÆs Mina Sur, to determine the effect on gas retention time when used in conjunction with drill cuttings in production shots with a blasthole diameter of 10 5/8ö. Longer retention times were expected to provide greater fragmentation and muckpile looseness. In order to determine which plug showed the greatest enhancement of stemming retention, 17 trials were carried out û in the first seven tests, only the Cats were used, in the following five tests, a combination of Cats, Taponex and Stemtite were used, and in the last three tests, only the Taponex plug was used. In the last two blasts, drill cuttings were used without stemming plugs as a benchmarking exercise. Blasts were filmed with a high-speed camera, which could record the detonation times of holes using rolls of shock tube tied to stakes at the hole collars, to act as flash indicators. The camera allowed measurement of the time between charge detonation and the commencement of stemming ejection. In order to obtain the statistical parameters of each one of the stemming plugs used in the 17 trials carried out, the Minitab software was used and a data base was created for it, corresponding to each stemming configuration with the retention time obtained from the filming. Once it was determined which stemming plug provided the greatest enhancement of stemming retention, the impact on the fragmentation was analysed using the WipFrag image analysis software to estimate the muckpile size distribution.

Jan 1, 2004

The Martha Mine is an open cut operation situated in the midst of Waihi town which has been operating since 1987. Since then it has produced some 700,000 ounces (oz) of gold and 4,000,000 oz of silver from an epithermal vein system. It was previously worked by underground methods from 1880 to 1952 to a depth of 600 m producing 5 million ounces (Moz) of gold and 30 Moz of silver. The licensed reserves are scheduled to be mined out early in the year 2001. An additional economic resource of over 750,000 oz of gold has been identified which can be accessed by deepening the pit a further 50 m along the main Martha vein. This would extend the mine life to the year 2007. Applications were lodged on 4 July 1997 for the necessary resource consents, discharge permits and variations for the current Mining Licence conditions. The extension project requires approval from the District Council, Regional Council and the Ministry of Commerce. The consent and permitting process is extensive because of the urban setting of the mine. An approach to keep the extension as similar as possible to the current operations was adopted to give the community and authority confidence in the proposed extension. This paper summarises the extension detail and the permitting requirements it has had to address despite the extension being "more of the same". The permitting process has been exhaustive with little recognition of the standards already achieved in protecting the environment which are recognised worldwide.

Jan 1, 1997

A BRIEF account of the water-power potentialities of Tasmania is given, together with possible applications of the power commercially. A short history of the Waddamana development is then given, and the main works are described.WATER POWER POTENTIALITIESTasmania is an island of approximately 24,500 sq. miles area (exclusive of small islands). An area of 12,000 sq. miles is at an elevation of over 1000 ft. above sea-level, and of 2300 sq. miles at over 3000 ft. above sea-level. Average annual rainfall varies from' over 140 in. on West Coast to 20 in. on East Coast, and for the whole island averages 48.5 in. The total water-power available, based on conservative figures for run off and efficiency, is 6,000,000 b.h.p. Only a small percentage of this total will, under present conditions, and under likely conditions in the immediate future, be capable of being developed commercially.An extensive hydro-metric and water-power survey is at present being carried out in order to collect information concerning the most promising water-power schemes. Already 400,000 b.h.p. has been located in different proposals, and there is still a large area of very likely country along the West Coast which has not been investigated yet. Tasmania thus has very vast power potentialities, which, combined with good ports and communications and wonderful mineral and timber resources, should augur well for the establishment of the great key industries, particularly chemical and metallurgical.

Jan 1, 1923

The Hodgkinson goldfield contains over 50 separate lodes which have produced 9.7 tonnes of gold from ore averaging 37 g/t. The gold-quartz.lodes are spatially separate from tin and base metal mineralisation, associated with Permo-Carboniferous granitoids in the province. Host rocks consist of sub-greenschist metamorphic grade shale, conglomerate, sandstone, spilite, chert and melange. Oreshoots average 100 x 30 x I m in Size and are contained in cross-cutting quartz. filled with steep-dipping shear zones. Mineralisation in the quartz. consists of gold with minor arsenopyrite, pyrite, galena, and sphalerite with rare calcite. Scheelite and stibnite are aSSOCIated locally but are not necessary proximal or coeval. Alteration is restricted to centimetre-scale phyllic selvedges adjacent to the veins and veinlets.Most of the lodes and vems occur within six main 1.5 x 0.25 km mineralised areas whIch may indIcate separate areas of fluid flow and precipitation. Ore controls differ m melange- and sandstone-hosted lodes but several features are common to all the oreshOOlS. Most of the lodes occur within zones of duauon that are caused by splaymg, refraction through carbonaceous shale-sandstone contacts, or by rolls and changes in strike and dip.Lithologic contacts may act as areas of permeability and chemistry contrast which focus fluid and precipitation. Tensional veins are common between parallel shear zones and curved spur veins occur both in the hangingwall and footwall of the main fissure. Intersections and contacts may have enhanced local fluid flow in many of the areas.

Jan 1, 1991

In a previous paper in these Proceedings(l) it was shown from a study of the bore'logs and analyses then available, that the brown coal seams of the Latrobe Valley areas show progressive changes in composition with increase of stratigraphic depth, equivalent to a rise in rank: These changes appear as:(i) An increase in ultimate carbon and hydrogen, calculated on a dry, ash-free basis, and(ii) An increase in calorific value, calculated on the, same basis, each of the order of 1% per 100 ft. of depth, and(iii) A decrease in moisture content of the order of 0.5% to 1.0% per 100 ft. of depth. The decrease in moisture content was found to be less regular than the other changes. In some bores the moisture content remained more or less constant throughout a single seam, or fluctuated slightly, decreasing stepwise on passage to the seam below.It was noted that individual analyses within a bore showed irregular fluctuations, tending to mask the general trend in composition change. This was particularly so for the moisture content, which had been determined for 10 ft. lengths of core, whereas the other components were determined for 50 ft. lengths of core. Averaging of the moisture contents over 50 ft. lengths reduced the masking effects of these irregularities. The suggestion was made that these irregularities arose from variations in the relative proportions of pollen, resin, wood (lignite) and woody attritus composing the core. Such variation in the coal seams has since been established in some detail for the...

Jan 1, 1948

By A Califice, E Pirard, M Benzaazoua, R Mermillod-Blondin

Image analysis is more and more considered as a powerful tool for quantitative analysis of ores and concentrates. Apart from being very much dependent upon the quality of the imaging principle and the adequacy of the imaging sensor, quantitative image analysis is also the result of a long chain of material, image and pixel sampling strategies. The most reliable way to demonstrate the accuracy and the precision of modal analysis obtained from images is probably to use a blended material with known proportions of pure minerals and confront the results obtained from different analytical tools. This operation has been performed on a series of mineral blends mimicking the tailings of a flotation operation in a plant operating on massive sulfide mineralisation. The intention was to be able to better predict acid mine drainage potential from direct mineralogical observations. Selected sulfide minerals: pyrite (Py), pyrrhotite (Po), chalcopyrite (Cp) and sphalerite (Sp) were mixed as variable amounts within gangue minerals (quartz, calcite, dolomite and siderite). Typical blends contained six to ten per cent (Py); one to three per cent (Po); 0.5 to one per cent (Cp, Sp); 15 to 40 per cent carbonates and 40 to 75 per cent quartz by mass. The 1000 gram blends were carefully sampled and prepared for X-ray diffraction, chemical analysis by ICP-AES and optical microscopy image analysis (OMIA). Results obtained for the different analytical techniques are compared and discussed revealing the limits and strength of each method in the accurate quantification of sulfide and gangue minerals.

Jan 1, 2008

By A. W. CAMERON

Over a 4 hr period of steady plant operation samples were taken at regular intervals from points within the grinding and flotation circuits. Size fractions of each composite sample were analysed for lead and zinc values and a comparison made of the behaviour of the lead, zinc, and gangue minerals, size by size throughout the whole circuit.Analysis of the breakage behaviour using dynamic mathematical models gave the rates of breakage of each size fraction of each main component. Simulated grinding experiments showed that the main grinding circuit had some reserve capacity.In the lead flotation circuit all lead values between 150 mesh (Tyler) and 5 microns had excellent flotation characteristics, and a major proportion had a high rate of flotation which was practically the same for all sizes. Increase in overall lead recovery would require improvement of the flotation behaviour of ultra-fine lead minerals. These were slow floating and were responsible for approximately half the lead losses in the final zinc concentrate and one-third of the lead losses in the final tails.All sizes of zinc values between 100 mesh and 7 microns floated well and the performance of the zinc flotation circuit was mainly limited by the relatively poor flotation of the very fine zinc values. Approximately 60 per cent of the zinc lost to the final tails lay within the minus 7 micron fraction.The overall flotation conditions were conducive to very good selective flotation of the major proportion of the values and limitations in metallurgical performance were mainly due to the properties of the ultra-fines.

Jan 1, 1971

By Anderson C. G

The design of crawler-mounted conveyor belts installed for the handling of brown coal in open cuts is described.INTRODUCTIONThe State Electricity Commission of Victoria operates two brown coal open cuts at Yallourn, Victoria. The larger of these cuts produces n million tons per year of brown coal containing 67 per cent. moisture, and the smaller, at Yallourn North, produces 1-1 million tons of coal, with a moisture content of 50 per cent. A further open cut at Morwell with projected output of 9-10 million tons per year is in the course of development. In the larger Yallourn open cut coal is excavated by means of conventional bucket chain and bucket wheel dredgers and transport is by electric locomotives and saddle bottom trucks running on 90 cm. gauge railway tracks. At Yallourn North, coal until recently has been largely won by 3i and 5 cu. yd. electric shovels and transported by large diesel-enginedmotor trucks. The difficulty of keeping these trucks in operation during the winter periods owing to the v.ery wet conditions and the high cost of maintenance of these vehicles led to the introduction of a moving face conveyor system.EXISTING CONVEYOR INSTALLATION AT YALLOURN NORTHA conveyor system employing a 36-in. belt running at a speed of 400 ft. per minute and having a nominal capacity of 425 tons per hour with 50-h.p. drive units, was selected as appropriate for the output of the open cut. The conveyor structure adopted was a standard type as used in under...

Jan 1, 1955

Some attempt to minimise the importance of the greenhouse phenomenon. They do so by observing that global temperatures are still within the range of recent variations and that it is still not possible to predict with any certainty what the precise consequences of greenhouse will be. Both observations are true though the conclusion is dangerous. Ice core samples from the Antarctic and Greenland have established the composition of the atmosphere and global temperature over almost the last 200,000 years. The profile of these twin variables shows that they have moved together; when the concentration of CO2 has been up, the temperature has been up and when the former has been down the temperature has also been down. Over almost 200,000 years the CO2 content has ranged between lows of 190 ppm and highs of 270ppm. In just 200 years of industrialisation the CO2 content of the atmosphere has increased to 350 ppm, 50% higher that the mean of the last 200,000 years and 30% higher than at the start of the industrial age. Carbon dioxide is now increasing at 0.5% p.a. Methane, about twenty times more effective than CO2 as a greenhouse gas has doubled in concentration in the last two centuries and is now increasing at about 1.1% p.a. Methane, nitrous oxide, chlorofluorocarbons together with a number of lesser gases will double the greenhouse effect due to in creased CO2 over the next 40-50 years.

Jan 1, 1989

Multiple regression analysis applied to ventilation records of stopes on the Witwatersrand showed that it was possible to estimate the temperature increase per 100 ft for a new stope with a standard deviation of 0' 5°F. The analysis also threw light on the relative effect on this estimate of five different factors.Application of the analysis of variance technique to a factorialexperiment on the production of dust during rock drilling showed that the effects of air and water pressure were quite small but also showed that the between-machine variation was unexpectedly large.Analysis of variance also proved a valuable means for assessing the bias and random error variation of aneroid barometers used in pressure surveys.INTRODUCTIONThree examples are discussed to illustrate the use of the techniques of "analysis of variance" and "multiple regression" in research on the ventilation of gold mines. Although the techniques are basically similar their application to each example has a slightly different emphasis and the examples were chosen for this reason. The variation ofemphasis is as follows.1. Temperature increases in stapes. This work was undertaken as a preliminary to possible extensive controlled experimentation (Lambrechts, 1959). Data collected in routine work by the mine ventilation personnel were used for analysis and no specific experiments were carried out. The main object was to see whether such data could provide any strong pointers to factors which might be worthy ofmore careful study and to factors which could initially be ignored. Of course, there was hope that the study might provide an empirical formula for rough forecasts.

Jan 1, 1962

By P S. R Reddy, V N. Misra

The iron ore deposits in India are rich in iron content but usually contains high alumina and silica. The current industrial practice results three products namely lump, fines and slimes. The lump and fines with or without beneficiation are directly fed to the blast furnace and sintering plant respectively while the super fines or slimes are presently discarded as tailings. Around eight million tonnes of iron ore slimes estimated as being 15 - 20 per cent by weight of run of mine ore containing ~48 - 62 per cent Fe, is discarded every year causing huge loss to the Indian iron ore industry. In addition, these slimes stored in massive tailing ponds pose enormous environmental and ecological problems. Extensive investigations have been carried out on three iron ore slimes generated at Barsua, Bolani and Meghahatuburu iron ore mines of India. The studies include physical, chemical, mineralogical characterisation and beneficiation studies adopting conventional hydroclyclone and wet high intensity magnetic separator. The characterisation studies have indicated that most of the gangue minerals are accumulated in finer fractions. The studies on hydrocyclone and magnetic separation had indicated that an iron concentrate of 61 - 65 per cent Fe could be obtained by recovering 59 - 78 per cent of the iron values present in the slimes. Depending on the nature of the slimes the alumina and silica levels have been brought down to 2.5 and 1.0 per cent respectively. The effects of different operating and design variables along with detailed characterisation studies have been highlighted.

Jan 1, 2002

By Jianmin H, Shuqin L

This paper analyses problems associated with cement based technique in metal mines, and provides a new mining technique of packing the tailing sand and water into solidified material. In the new technique, the cementing agent is a solidifying material containing a high percentage of water and tailing sand. The sand and water solidify into the packed body at the same time. The slurry of the packed body varies widely in concentration and suitability for different mining conditions. It consists of two liquid materials, each composed of a cementing agent, tailing sand and water. The thick liquids are transported to the packing place at the same time. Their pumping property is good; they do not solidify or block up the pipe during transportation. At the packing place, the two liquids are evenly mixed.The mixed liquid flattens out well and its solidifying speed is so fast that it solidifies into a packed body within a very short period of time. The compressed strength of the packed body can reach 1.2-5MPa in one day, which satisfies the requirements of different mining conditions. By using the new mining technique, all the tailing sand can be used again in the packed body and it is low in cost and investment. The mechanisation is easy and thus the working efficiency is increased and the use of labour decreased. Contamination problems caused by i. Dr & Prof. Beijing Graduate School, China University of Mining Science and Technology,Beijing,China 2 Northern .Jiaotong University, Beijing,China 3. Beijing Graduate School, China University of Mining, Science and Technology

Jan 1, 1992

By Edwards A. B

The occurrence of gold so finely divided as to be invisible under the microscope, even at magnificatioIl; of 1,200 to 2,000 diameters, is rare, but not unknow;n.. Burg (1930) has described an occurrence of such "invisible gold" from the 25/37 vein of the Bradisor mine at Brad in the Siebengeburgen district of Rumania (Transylvania), in which the pyrite contained up to 150 grams of gold per ton (approx. 5 oz. gold per ton). Hoffman (1931) has described an epithermal silver gold ore from the Titiribi district of Columbia (South America), in which "invisible gold" to the extent of 150 grams per ton occurred in the pyrite, arsenopyrite and zincblende; and Haycock (1937) has recorded it in pyrite from the Surf Point mine (Canada), which contained no visible gold, but yielded positive testsfor gold whenever samples drilled from polished surfaces of apparently homogeneous pyrite were tested spectrographically.The examination of a cyanided pyritic concentrate, together with a sample of the ore from which it' was prepared, from the Dolphin East lode at Vatukoula, Fiji, carried out as part of the programme of the Mineragraphic Section of the Council for Scientific and Industrial Research, Australia, has established a further occurrence of "invisible gold" in pyrite and arsenopyrite. The Dolphin pyrite contains about 25 oz. gold per ton, and the arsenopyrite contains more than 33 oz. per ton.The concentrate has been prepared for the Mill Superintendent of the Emperor Gold Mining Co. by L. Mayger...

Jan 1, 1946

By S Cossfo Torres

Minera PenmontÆs gold mine, La Herradura, is located at the northwestern Mexican state of Sonora and consists of an open pit mine and heap leaching facility for the recovery of gold. The production began in July 1998, and today La Herradura produces about 200 000 oz of gold per year. Standard sampling and assaying protocols were established during exploration based on heterogeneity tests. However, at deeper zones of the pit, coarse gold became evident and it was necessary to develop a new sampling protocol in order to minimise the fundamental error and improve the metallurgical accounting. This new sampling protocol requires cyanide leaching the whole pulverised sample from the blasthole, assaying pregnant solution and fire assaying 50 g of tailings. A sampling study was performed on 195 blasthole samples analysing gold under the standard and new protocols. The new protocol demonstrated that the original one underestimated gold grade by approximately 27 per cent, with impacts on ore misclassification and metallurgical accounting. According to the ore grade control system, mineral with gold grade above 0.35 g/t was sent to the leach pad and mineral with lower grade was classified as waste. Based on the sampling study, it was decided to classify a low-grade ore with gold assay between 0.20 and 0.35 g/t which was stored for future treatment under better gold prices or should mine production decrease. These two conditions were present during 2005 and 2006, when about 3.5 Mt of low-grade ore were leached.

Jan 1, 2008

Titanium tetrachloride may be reduced to trichloride with hydrogen on a hot tungsten filament and other hot: surfaces. The effect of temperature, time and size of t.he filament on the rate of reaction has been studied. A laboratory unit for preparing a few pounds of trichloride at a time is described. The yield of trichloride is 2·2 - 2·4 gm/sq. cm. of hot surface/hour at 1,000°C. and the energy requirement 3'5-4 kWh/pound of trichloride produced.'INTRODUCTIONThe lower chlorides of titanium (titanium tri- and dichlorides) are amenable to electrolysis, being electrovalent in nature. Furthermore, relative to other oxygen-free titanium salts, they are easy to prepare in a pure state and their use has been investigated to prepare titanium. The experimentaJ work described in this paper deals with the preparation of titanium trichloride by reduction of titanium tetrachloride with hydrogen on hot surfaces.REDUCING AGENTS FOR TITANIUM TETRACHLORIDEThe free energy diagram (Fig. 1) for the formation of chlorides suggests that titanium' tetrachloride may be reduced to trichloride by various reducing agents such as aluminium, zinc, arsenic, antimony, tin, mercury, silver, hydrogen, sodium, potassium and magnesium. Ruff and Neumann (1923), Thorpe (1885) , Billy and Brasseur (1935) and Seurbert and Schmidt (1892) used various metals to reduce tetrachloride to trichloride and found that the yields were invariably low and the products difficult to separate.Besides the metallic reducing agents, hydrogen may also be used to reduce titanium tetrachloride to trichloride.Goerges and Stahler (1909) obtained a deposit of violet...

Jan 1, 1956

By M Freeman

Significant improvements in the metallurgical performance at Pasminco Rosebery Mine have occurred over the last two years. Lead and silver recoveries have increased by more than ten per cent, lead concentrate grades are four per cent higher, zinc recoveries are up 1.5 per cent, and zinc concentrate grades have increased by two per cent. These impressive results have not occurred through the application of some new wonder reagent, but through the application of some fundamental metallurgical principals. The basic premise applied during this study was that to make an effective separation you must first achieve liberation. Therefore, the laboratory and plant studies examined the liberation characteristics within the lead circuit, and the impact of regrinding. The work demonstrated clearly that galena/sphalerite liberation was poor, and significantly impaired plant performance. Further, it was suggested that the application of lead regrinding of the lead scavenger concentrate and the lead first cleaner tailing would not only benefit lead metallurgy, but also improve sphalerite flotation. A similar approach was then applied to the zinc circuit. From these studies, solutions applying simple circuit changes utilising existing equipment were proposed and implemented. Further to the improvements in galena/sphalerite liberation, significant improvements in concentrate quality were achieved by reducing variation within the process. The implementation of changes in work practices negotiated in the Enterprise Agreement, employee contributions and specialist training also made a significant contribution to the unprecedented improvement in the lead and zinc flotation circuit performance. The results of these plant changes are discussed, along with the suggested scope for further improvement.

Jan 1, 2000