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An overview is presented of water management practices in the mining industry over the last two decades, and examples are given of recently adopted water treatment practices. Areas highlighted include: the relationships between mines and the surrounding communities concerning water supplies; water balances at mines; ways of reducing water consumption through the use of dry tailings disposal and recycling; and the use of alternative sources of water supply, such as hyper-saline groundwater, desalinated water and seawater, and relevant treatment methods., An overview is presented of water management practices in the mining industry over the last two decades, and examples are given of recently adopted water treatment practices. Areas highlighted include: the relationships between mines and the surrounding communities concerning water supplies; water balances at mines; ways of reducing water consumption through the use of dry tailings disposal and recycling; and the use of alternative sources of water supply, such as hyper-saline groundwater, desalinated water and seawater, and relevant treatment methods.

Laboratory tests and plant trials were carried out on the processing of Au ore containing carbonaceous shale from the Jundee operation in Western Australia. Samples were taken from stockpiles and portions of the underground orebody which contain naturally occurring graphitic carbon, resulting in mild to moderately severe preg-robbing behaviour. The results showed the advantage of using carbon in leach compared with direct cyanide leaching and carbon in pulp. Addition of kerosene further enhanced recoveries for all circuit options. Factors which increased overall Au recoveries included keeping preg-robbing ore separate from non- graphitic ore, maximising gravity recovery, ensuring the plant operated in pure CIL mode, with no leaching prior to the first adsorption tank, ensuring activated carbon in the circuit remained active with kerosene addition and optimisation of reagent concentrations and residence times., Laboratory tests and plant trials were carried out on the processing of Au ore containing carbonaceous shale from the Jundee operation in Western Australia. Samples were taken from stockpiles and portions of the underground orebody which contain naturally occurring graphitic carbon, resulting in mild to moderately severe preg-robbing behaviour. The results showed the advantage of using carbon in leach compared with direct cyanide leaching and carbon in pulp. Addition of kerosene further enhanced recoveries for all circuit options. Factors which increased overall Au recoveries included keeping preg-robbing ore separate from non- graphitic ore, maximising gravity recovery, ensuring the plant operated in pure CIL mode, with no leaching prior to the first adsorption tank, ensuring activated carbon in the circuit remained active with kerosene addition and optimisation of reagent concentrations and residence times.

The experience is reviewed that was gained in the treatment of both clayey oxide Au ores and the supergene Cu-Au ores at the Boddington mine in Western Australia during the period 1989 to 2001. The development is described of a process flow sheet to allow economic treatment of the large low-grade competent Au-Cu resource underlying the oxide and supergene Au resource which resulted in the approval in 2006 of the 35 000 000 tpa Boddington expansion project. The flow sheet provides the basis of an energy-efficient comminution circuit including high-pressure grinding rolls and a treatment circuit with gravity separation and flash and conventional flotation to produce a precious metal low-grade Cu concentrate, followed by cyanide leaching of the rougher-scavenger and cleaner tailings streams. The circuit allows maximum removal of Cu minerals prior to cyanide leaching., The experience is reviewed that was gained in the treatment of both clayey oxide Au ores and the supergene Cu-Au ores at the Boddington mine in Western Australia during the period 1989 to 2001. The development is described of a process flow sheet to allow economic treatment of the large low-grade competent Au-Cu resource underlying the oxide and supergene Au resource which resulted in the approval in 2006 of the 35 000 000 tpa Boddington expansion project. The flow sheet provides the basis of an energy-efficient comminution circuit including high-pressure grinding rolls and a treatment circuit with gravity separation and flash and conventional flotation to produce a precious metal low-grade Cu concentrate, followed by cyanide leaching of the rougher-scavenger and cleaner tailings streams. The circuit allows maximum removal of Cu minerals prior to cyanide leaching.

The effects were investigated of bowl angle, bowl speed and feed rate on the recovery of very fine Au in leach tailings using a Falcon B separator. Best recovery of smaller than 20 microns was achieved with a 14 degree bowl. Replicate tests suggested that bowl speeds giving greater than 250 G do not offer additional Au or sulphide recovery for these feed materials. Extending the size analysis to smaller than 11 microns showed that the separator concentrated Au-bearing silicate particles down to 11 microns. The optimum recovery of sulphur was achieved with a 10 degree bowl. As the density of the fine Au-bearing particles is likely to be only just greater than gangue density, due to the small amounts of Au present, the optimum conditions for denser sulphide mineral recovery are an important indicator of the ability of the separator to recover very fine dense particles. Mass recovery to the concentrate was a function of bowl angle, with the greatest mass recovery occurring using the smallest angle. Mass pull was similar irrespective of amount of feed. When the feed mass was increased to 5 kg, sulphur recovery dropped significantly due to bowl overloading and Au recovery dropped about 10% compared with the lower feed-mass tests. Substantial recovery of Au in all size fractions was possible from the much coarser feed due to the higher proportion of free Au particles. Coarser particles in the fine leach feed may contain less gold, and the recovery of these particles is consequently lower., The effects were investigated of bowl angle, bowl speed and feed rate on the recovery of very fine Au in leach tailings using a Falcon B separator. Best recovery of smaller than 20 microns was achieved with a 14 degree bowl. Replicate tests suggested that bowl speeds giving greater than 250 G do not offer additional Au or sulphide recovery for these feed materials. Extending the size analysis to smaller than 11 microns showed that the separator concentrated Au-bearing silicate particles down to 11 microns. The optimum recovery of sulphur was achieved with a 10 degree bowl. As the density of the fine Au-bearing particles is likely to be only just greater than gangue density, due to the small amounts of Au present, the optimum conditions for denser sulphide mineral recovery are an important indicator of the ability of the separator to recover very fine dense particles. Mass recovery to the concentrate was a function of bowl angle, with the greatest mass recovery occurring using the smallest angle. Mass pull was similar irrespective of amount of feed. When the feed mass was increased to 5 kg, sulphur recovery dropped significantly due to bowl overloading and Au recovery dropped about 10% compared with the lower feed-mass tests. Substantial recovery of Au in all size fractions was possible from the much coarser feed due to the higher proportion of free Au particles. Coarser particles in the fine leach feed may contain less gold, and the recovery of these particles is consequently lower.

Two methods were used to investigate the relationship between the flotation of chalcopyrite and pyrite from an ore containing the two minerals. The first was to remove the xanthate species with an organic solvent and use ultraviolet spectroscopy to determine the concentration of dissolved species. The second involved determining xanthate decomposition in solution by measuring the residual xanthate concentration using ultraviolet spectroscopy. In both cases pH and temperature were varied and two collectors, sodium ethyl xanthate and potassium amyl xanthate, were used. The detection of heavy metal xanthates and dixanthogen on the surface of the sulphides after flotation and in the float tail water was used to relate changes in flotation recovery and grade to changes in the adsorbed species on the sulphides. The relationships could be applied to industrial flotation systems and allow for more accurate determination of optimal xanthate dosage and hence save on reagent costs., Two methods were used to investigate the relationship between the flotation of chalcopyrite and pyrite from an ore containing the two minerals. The first was to remove the xanthate species with an organic solvent and use ultraviolet spectroscopy to determine the concentration of dissolved species. The second involved determining xanthate decomposition in solution by measuring the residual xanthate concentration using ultraviolet spectroscopy. In both cases pH and temperature were varied and two collectors, sodium ethyl xanthate and potassium amyl xanthate, were used. The detection of heavy metal xanthates and dixanthogen on the surface of the sulphides after flotation and in the float tail water was used to relate changes in flotation recovery and grade to changes in the adsorbed species on the sulphides. The relationships could be applied to industrial flotation systems and allow for more accurate determination of optimal xanthate dosage and hence save on reagent costs.

Plant upgrades are described that were carried out in the Cadia Hill, New South Wales, circuit in April 2003 to improve concentrate quality and metal recoveries from the flotation circuit. The major changes involved circuit reconfiguration to allow all concentrates to be reground to less than 38 micrometres prior to final cleaning, increasing ball charge in the regrind mill to utilise fully all of the installed power, desliming of the cleaner scavenger concentrate to improve the circuit water balance and the inclusion of a gravity concentrator in the regrind circuit to recover more free gold to bullion. Once the flotation circuit had stabilised, concentrate launder modifications and increased froth crowding on the scavenger flotation cells permitted the introduction of stronger collectors without compromising concentrate quality or circuit stability. After 12 months, the Cu concentrate grade had increased by 2% and metal recoveries by 3% and 7% for Cu and Au, respectively. The project capital outlay has been repaid in less than three months., Plant upgrades are described that were carried out in the Cadia Hill, New South Wales, circuit in April 2003 to improve concentrate quality and metal recoveries from the flotation circuit. The major changes involved circuit reconfiguration to allow all concentrates to be reground to less than 38 micrometres prior to final cleaning, increasing ball charge in the regrind mill to utilise fully all of the installed power, desliming of the cleaner scavenger concentrate to improve the circuit water balance and the inclusion of a gravity concentrator in the regrind circuit to recover more free gold to bullion. Once the flotation circuit had stabilised, concentrate launder modifications and increased froth crowding on the scavenger flotation cells permitted the introduction of stronger collectors without compromising concentrate quality or circuit stability. After 12 months, the Cu concentrate grade had increased by 2% and metal recoveries by 3% and 7% for Cu and Au, respectively. The project capital outlay has been repaid in less than three months.

The use is discussed of computer simulation, based on information derived from laboratory batch test-work or plant and laboratory data, to provide an initial assessment of the benefits of a flash flotation circuit. Details are presented of an extension to a model that allows flash flotation performance to be obtained within a closed-circuit grinding operation. The extension incorporates the interaction of flash flotation on the downstream conventional flotation circuit and allows 'what if' scenarios to be explored. A description is given of the laboratory batch test methodology, carried out on a sample of Au ore, and the data derived from the tests which were used in the models and simulation. The results showed the effect of increasing the split to the flash cell on concentrate grade and recovery., The use is discussed of computer simulation, based on information derived from laboratory batch test-work or plant and laboratory data, to provide an initial assessment of the benefits of a flash flotation circuit. Details are presented of an extension to a model that allows flash flotation performance to be obtained within a closed-circuit grinding operation. The extension incorporates the interaction of flash flotation on the downstream conventional flotation circuit and allows 'what if' scenarios to be explored. A description is given of the laboratory batch test methodology, carried out on a sample of Au ore, and the data derived from the tests which were used in the models and simulation. The results showed the effect of increasing the split to the flash cell on concentrate grade and recovery.

The main body of flotation research is still carried out at university level but there has been a shift from smaller sponsored projects to large multi-client projects, with sufficient funding to tackle more complex flotation problems, or to the formation of collaborative research groups or research centres. Flotation research is performed at a more focused level by vendors of flotation reagents and flotation equipment. Some mining companies may take on in-house research projects specifically to solve metallurgical problems at a mine site or a new deposit, sometimes requiring input from university specialists. Modelling tools are likely to become more prevalent to solve the complex problems of the future., The main body of flotation research is still carried out at university level but there has been a shift from smaller sponsored projects to large multi-client projects, with sufficient funding to tackle more complex flotation problems, or to the formation of collaborative research groups or research centres. Flotation research is performed at a more focused level by vendors of flotation reagents and flotation equipment. Some mining companies may take on in-house research projects specifically to solve metallurgical problems at a mine site or a new deposit, sometimes requiring input from university specialists. Modelling tools are likely to become more prevalent to solve the complex problems of the future.

The application of high-pressure grinding rolls (HPGR) at pilot and industrial scale in Australia is described. The Argyle diamond mines introduced the first of two smooth segmented rollers in 1990, and a studded roll HPGR in 2002. Pilot plant trials were conducted by the Kalgoorlie Consolidated gold mine in 1993, and the same unit was used in trials at the Boddington gold mine in 1996. Laboratory and pilot-plant facilities in Australia are described, together with the JKSimMet software package for the modelling and simulation of HPGR performance and an HPGR wear model., The application of high-pressure grinding rolls (HPGR) at pilot and industrial scale in Australia is described. The Argyle diamond mines introduced the first of two smooth segmented rollers in 1990, and a studded roll HPGR in 2002. Pilot plant trials were conducted by the Kalgoorlie Consolidated gold mine in 1993, and the same unit was used in trials at the Boddington gold mine in 1996. Laboratory and pilot-plant facilities in Australia are described, together with the JKSimMet software package for the modelling and simulation of HPGR performance and an HPGR wear model.

Although very high Cu and Au recoveries are possible from the Western Australian bulk concentrates, which typically contain 8% chalcopyrite, 8% chalcocite, 45% pyrite and 80-120 g/t gold, the cost of lime for neutralisation and power for O2 generation makes conventional high- temperature pressure oxidation with solvent extraction, electrowinning and cyanidation of the leach residue uneconomic. A series of batch and locked-cycle testwork programmes, supported by Metsim modelling, established that the same high overall recoveries can be obtained by two- stage split leaching of chalcocite at 90-110 degrees C and chalcopyrite at 220-250 degrees C. Advantages of this split circuit include minimising the required capacity of the high-temperature autoclave, reducing lime consumption and simplifying process water balance. The integration of the leaching stages into the overall circuit is discussed., Although very high Cu and Au recoveries are possible from the Western Australian bulk concentrates, which typically contain 8% chalcopyrite, 8% chalcocite, 45% pyrite and 80-120 g/t gold, the cost of lime for neutralisation and power for O2 generation makes conventional high- temperature pressure oxidation with solvent extraction, electrowinning and cyanidation of the leach residue uneconomic. A series of batch and locked-cycle testwork programmes, supported by Metsim modelling, established that the same high overall recoveries can be obtained by two- stage split leaching of chalcocite at 90-110 degrees C and chalcopyrite at 220-250 degrees C. Advantages of this split circuit include minimising the required capacity of the high-temperature autoclave, reducing lime consumption and simplifying process water balance. The integration of the leaching stages into the overall circuit is discussed.

Details are given of an investigation into potential water sources and the effect of water quality on the flotation of the Cu-Au bearing ore which was carried out prior to commissioning of a new concentrator at the Cadia Hill gold mine in New South Wales. An extensive water quality survey of all major potential water supplies to the concentrator was undertaken and the flotation performance in each of the waters was investigated on a laboratory scale. The results showed that water quality could have a significant influence on the flotation process. Since operations began, water quality surveys of both make-up water sources and recycled flotation circuit waters have been carried out on a regular basis. A water quality database has been established which can be compared with plant performance. A series of laboratory flotation tests were conducted using a standard ore and waters from various streams to monitor changes in flotation performance as a function of water quality., Details are given of an investigation into potential water sources and the effect of water quality on the flotation of the Cu-Au bearing ore which was carried out prior to commissioning of a new concentrator at the Cadia Hill gold mine in New South Wales. An extensive water quality survey of all major potential water supplies to the concentrator was undertaken and the flotation performance in each of the waters was investigated on a laboratory scale. The results showed that water quality could have a significant influence on the flotation process. Since operations began, water quality surveys of both make-up water sources and recycled flotation circuit waters have been carried out on a regular basis. A water quality database has been established which can be compared with plant performance. A series of laboratory flotation tests were conducted using a standard ore and waters from various streams to monitor changes in flotation performance as a function of water quality.

The mine reserve consists of a diverse range of Cu and Au mineralisation with relatively high levels of gravity-recoverable Au. The complexity of the orebody required a flexible plant to recover Cu and Au values by a combination of gravity recovery, flotation and concentrate leaching. The treatment of ore from a high-grade underground mine and a variable-grade open pit required a dual-train approach from primary crushing to concentrate production in order to minimise the requirement for different ore types. An extensive sampling, testing and technical evaluation programme was undertaken to identify processing characteristics and the optimum process route for each ore type. The design experience included the development of robust primary stockpiling and pebble crushing concepts, detailed review of scale-up of SAG and ball-mill pilot plant data to plant scale, the selection of two 15 MW SAG mills and two 13 MW ball mills for the grinding circuit and detailed modelling of the pyrite concentrate carbon-in-leach (CIL)/sulphidisation acidification recovery thickening (SART)/acidification volatilisation recovery (AVR) process to optimise design for Cu and cyanide recovery from leach liquors., The mine reserve consists of a diverse range of Cu and Au mineralisation with relatively high levels of gravity-recoverable Au. The complexity of the orebody required a flexible plant to recover Cu and Au values by a combination of gravity recovery, flotation and concentrate leaching. The treatment of ore from a high-grade underground mine and a variable-grade open pit required a dual-train approach from primary crushing to concentrate production in order to minimise the requirement for different ore types. An extensive sampling, testing and technical evaluation programme was undertaken to identify processing characteristics and the optimum process route for each ore type. The design experience included the development of robust primary stockpiling and pebble crushing concepts, detailed review of scale-up of SAG and ball-mill pilot plant data to plant scale, the selection of two 15 MW SAG mills and two 13 MW ball mills for the grinding circuit and detailed modelling of the pyrite concentrate carbon-in-leach (CIL)/sulphidisation acidification recovery thickening (SART)/acidification volatilisation recovery (AVR) process to optimise design for Cu and cyanide recovery from leach liquors.

The semiautogenous grinding mill at Cadia Hill, in New South Wales, is 12 m in diameter and was installed after two years of extensive laboratory and pilot-plant testwork. The hard, competent ores required the use of a both variable-speed SAG mill drive and pebble crushing. Grinding power requirements were based on the use of in-house software, pilot-plant data and JKSimMet simulations. Surveys during the first four weeks of operation showed that the feed size distribution, transfer rate and pebble rate were less than design levels, so the SAG ball load was increased and total comminution circuit power consumption rose to 0.5 kWh/t higher than predicted; the JKSimMet model was modified to take account of the variations., The semiautogenous grinding mill at Cadia Hill, in New South Wales, is 12 m in diameter and was installed after two years of extensive laboratory and pilot-plant testwork. The hard, competent ores required the use of a both variable-speed SAG mill drive and pebble crushing. Grinding power requirements were based on the use of in-house software, pilot-plant data and JKSimMet simulations. Surveys during the first four weeks of operation showed that the feed size distribution, transfer rate and pebble rate were less than design levels, so the SAG ball load was increased and total comminution circuit power consumption rose to 0.5 kWh/t higher than predicted; the JKSimMet model was modified to take account of the variations.

The need for quantitative image analysis to determine mineral liberation and texture led to the development of quantitative evaluation of minerals using scanning electron microscopy (QEM.SEM). Particulate materials are cast into sample blocks with crushed graphite and chlorinated epoxy resin, so that random orientations and sections are exposed and well separated. Each particle is distinguished by its back-scattered electron contrast and its minerals classified mainly from the X-ray spectrum. The much faster Philips XL40 SEM provides a unique large image which allows accurate identification of minerals from back-scattered electron intensity, using the X-ray spectrum only to identify ambiguous pairs. An updated version of the QEM.SEM is the Leo QEMSCAN system. At New Celebration gold mine in Western Australia, the combination of QEM.SEM analysis with grinding plant simulation allowed both treatment rates and recovery to be improved., The need for quantitative image analysis to determine mineral liberation and texture led to the development of quantitative evaluation of minerals using scanning electron microscopy (QEM.SEM). Particulate materials are cast into sample blocks with crushed graphite and chlorinated epoxy resin, so that random orientations and sections are exposed and well separated. Each particle is distinguished by its back-scattered electron contrast and its minerals classified mainly from the X-ray spectrum. The much faster Philips XL40 SEM provides a unique large image which allows accurate identification of minerals from back-scattered electron intensity, using the X-ray spectrum only to identify ambiguous pairs. An updated version of the QEM.SEM is the Leo QEMSCAN system. At New Celebration gold mine in Western Australia, the combination of QEM.SEM analysis with grinding plant simulation allowed both treatment rates and recovery to be improved.

Dump leaching at Telfer in its present form began towards the end of 1988. Initial plans called for a dedicated dump leach facility capable of processing ore at an annual rate of 2 000 000 tonnes. Construction began in mid-1988 of a dump leach pad, collection and process ponds, a carbon adsorption system, and reagent addition and ancillary facilities. During construction additional reserves were proven and design throughput was increased to 4 000 000 tonnes/a. Leach pad capacity and solution supply systems were upgraded to handle the larger throughput. Space limitations meant that the carbon adsorption/elution facilities were not upgraded. In 1990 it was decided to expand the operation to 6 000 000 tonnes/a and construction of a second leach pad began in 1991. A further two pads have been constructed, the last in 1994 bringing total annual throughput to 10 000 000 tonnes., Dump leaching at Telfer in its present form began towards the end of 1988. Initial plans called for a dedicated dump leach facility capable of processing ore at an annual rate of 2 000 000 tonnes. Construction began in mid-1988 of a dump leach pad, collection and process ponds, a carbon adsorption system, and reagent addition and ancillary facilities. During construction additional reserves were proven and design throughput was increased to 4 000 000 tonnes/a. Leach pad capacity and solution supply systems were upgraded to handle the larger throughput. Space limitations meant that the carbon adsorption/elution facilities were not upgraded. In 1990 it was decided to expand the operation to 6 000 000 tonnes/a and construction of a second leach pad began in 1991. A further two pads have been constructed, the last in 1994 bringing total annual throughput to 10 000 000 tonnes.

The installation of the plant has demonstrated that sulphide-bearing gold cyanidation tailings can be economically exploited, though the viability of the process depends largely on the combined use of low-cost gravity separation and an energy-efficient fine grinding circuit, as well as favourable mineralogy. Gravity gold recovery in the tailings treatment plant has been lower than expected, due to differences in ore characteristics and residue contents and to excessive clay fines in the Reichert cone feed. Changing the cut-off point of the desliming cyclones would improve gravity performance if additional dilution water was available for the gravity circuit; for a similar project with a longer life a water recovery circuit would probably prove viable. Operating costs have been 30% lower than anticipated, partly offsetting the revenue lost to low residue grades and gold recoveries. Payback period has been extended from 13 to 23 months., The installation of the plant has demonstrated that sulphide-bearing gold cyanidation tailings can be economically exploited, though the viability of the process depends largely on the combined use of low-cost gravity separation and an energy-efficient fine grinding circuit, as well as favourable mineralogy. Gravity gold recovery in the tailings treatment plant has been lower than expected, due to differences in ore characteristics and residue contents and to excessive clay fines in the Reichert cone feed. Changing the cut-off point of the desliming cyclones would improve gravity performance if additional dilution water was available for the gravity circuit; for a similar project with a longer life a water recovery circuit would probably prove viable. Operating costs have been 30% lower than anticipated, partly offsetting the revenue lost to low residue grades and gold recoveries. Payback period has been extended from 13 to 23 months.

An understanding of the behaviour of liberated and unliberated minerals in comminution and classification circuits is important in ensuring that maximum liberation occurs without unnecessary and costly overgrinding. A study of liberation performance was carried out on Newcrest Mining's New Celebration grinding circuit. Data from circuit surveys and QEM SEM analyses of individual streams were used to obtain a detailed picture of the liberation and classification of gold-associated pyrite. The relationship between the QEM SEM analyses and results from laboratory leach tests are also discussed. Mathematical models were constructed of the performance of the ball mills and cyclones and simulations were carried out to investigate ways of improving the circuit performance., An understanding of the behaviour of liberated and unliberated minerals in comminution and classification circuits is important in ensuring that maximum liberation occurs without unnecessary and costly overgrinding. A study of liberation performance was carried out on Newcrest Mining's New Celebration grinding circuit. Data from circuit surveys and QEM SEM analyses of individual streams were used to obtain a detailed picture of the liberation and classification of gold-associated pyrite. The relationship between the QEM SEM analyses and results from laboratory leach tests are also discussed. Mathematical models were constructed of the performance of the ball mills and cyclones and simulations were carried out to investigate ways of improving the circuit performance.

Grade control is an important factor in the success of a mining operation. Grade control criteria vary with the commodity being mined and financial considerations determine the actual grade control criteria at any one mine site. Mining and metallurgical capital and operating costs are ultimately used to ascertain the cut-off grades used for grade control. The metallurgical characteristics of the ore and, ultimately, the equipment selection and plant design are based on samples provided by the exploration geologist. Ore variability will impact on metallurgical performance and hence on the economics of a mining operation. Therefore, good liaison between the grade control geologist and the metallurgist should lead to a more consistent and profitable operation. The metallurgical considerations required for a gold processing operation and the practical operating problems associated with each unit process are discussed., Grade control is an important factor in the success of a mining operation. Grade control criteria vary with the commodity being mined and financial considerations determine the actual grade control criteria at any one mine site. Mining and metallurgical capital and operating costs are ultimately used to ascertain the cut-off grades used for grade control. The metallurgical characteristics of the ore and, ultimately, the equipment selection and plant design are based on samples provided by the exploration geologist. Ore variability will impact on metallurgical performance and hence on the economics of a mining operation. Therefore, good liaison between the grade control geologist and the metallurgist should lead to a more consistent and profitable operation. The metallurgical considerations required for a gold processing operation and the practical operating problems associated with each unit process are discussed.