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By R Berry

This paper looks at two ways to estimate the bulk mineralogy of the rocks for assay intervals. The aim is to find an efficient indicator of the most common minerals in the rock. Phase (modal) analysis has traditionally been done using visual methods such as point counting and image analysis. A modern version of this process is the X-ray point counting routine using the SEM-EDS based software. These methods are too slow and expensive for routine analysis of bulk sample mineralogy at the normal assay spacing.Two sources of data were considered that provide information that can be used to determine the mineral abundance in assay samples. The most widely applied method is (semi-) quantitative X-ray diffraction (QXRD). The QXRD method is most applicable to major minerals and has limited application to minerals at low abundance. The nominal detection limit is 0.5 per cent. Values below five per cent have large errors. A second, less common, method is calculation of mineralogy from chemical assay data. Conversion of chemical analyses to mineralogical analyses depends on the unique chemical composition of each mineral. Elements only found in one mineral are easily accounted for, but many compositions are ambiguous. Deciding on the actual mineralogy is not simple. Recalculation of mineral mode from chemical analyses is more accurate than QXRD when the correct minerals, and mineral compositions, are known.Where only a few QXRD analyses are available they can be used to setup a protocol for calculation of mineralogy from assay data. Linear programming works well in this environment. The best results are obtained when both H2O and CO2 are directly measured. Loss-on-ignition (LOI) should be included if these are not available.Where both QXRD and chemical analysis are available for all samples, the best results are obtained using the least squared method to merge the data sets assuming QXRD has much higher analytical errors than chemical assays. The combined method provides more robust results because the high abundance minerals are controlled by the QXRD measurements while the chemical assays improve the precision for low abundance minerals

Sep 5, 2011

The Youga gold mine in Burkina Faso, West Africa, is scheduled to produce first gold in the third quarter of calendar 2007. The project is expected to produce 540 000 oz of gold over 6.5 years, from a conventional open pit carbon in leach mine. The direct capital cost of the project is US$47 M. Etruscan Resources Inc, a Canadian based gold exploration and development company, holds 90 per cent of the project company Burkina Mining Company SA (BMC). The state of Burkina Faso is free-carried for the remaining ten per cent of BMC. RMB Resources Limited, of Melbourne Australia, has arranged a US$35 M project finance facility for BMC (the Borrower) as the majority of funding for the Youga project development. The loan facility was provided by RMB Australia Holdings and Macquarie Bank Ltd (the Lenders) each as to 50 per cent in the lending syndicate. The Lenders also provided a gold hedging facility to guarantee a minimum price for the projectÆs gold production. Burkina Faso is a francophone country, part of the Economic and Monetary Union of West Africa (UEMOA). The project financing is the first such facility to be arranged for a mining project in Burkina Faso, with little or no precedent available in other UEMOA countries. Therefore, while the Youga gold project is forecast to be a robust project applying conventional mining and processing methods, a number of legal, technical, commercial and social issues required identification, evaluation and resolution before the loan could be established. This paper describes that evaluation process and the resolution of these issues. In-house experience and expertise in both the LendersÆ and BorrowerÆs teams has facilitated a greater understanding and æownershipÆ of the issues and hence provided comfort with risks which might not normally be taken in conventional project developments and project financings. As a result, blanket external reviews were not required by the Lenders. A long-standing relationship between Etruscan and RMB Resources also facilitated solutions to a number of issues. Nevertheless, a degree of flexibility by the Lenders, the Borrower and Etruscan was required to successfully close the financing.

Jan 1, 2007

By Shuguang L, Jing S

Computer controlled truck dispatching systems is a newly-developed technical system in surface mines with shovel-truck technology. The real time dispatching criteria, as the core of optimal truck dispatching software, are analysed and a new criterion, complex dispatching criterion, is put forward. Dispatching criteria previously proposed include: 1. `Earliest Loading' (EL) - to assign a truck to the shovel at which the truck can be loaded earliest. 2. `Max Truck' (MT) - to assign a truck to the shovel at which the truck will have the shortest waiting time. 3. `Max Shovel' (MS) - to assign a truck to the shovel which has been waiting for the longest time. 4. `Minimum Saturation Degree' (MSD) - to assign a truck to the shovel at the branch where the truck flow saturation degree is minimum. 5. `Minimum Saturation Degree with Truck Cycle time consideration' (MSDTC) - as the MSD principle, with the truck travel cycle also considered. 6. `Fixed Dispatching' (FIX) - to assign trucks to fixed shovels. A system simulation model has been set up for truck dispatching system analysis. A few surface coal mines have been selected for case studies. Equipment waiting times and shift production of the system are obtained as the main outputs of the model. Simulation results show that each dispatching criterion has its own applicable range, no one can absolutely dominate the others. For example, when under-trucked, EL and MT criterion can obtain higher production than the MS criterion; when over-trucked, it is vice versa; while for MSD criteria, better results can be obtained no matter how the truck numbers fluctuate. From the view point of shovel output balance, large fluctuations occur for EL and MT criteria, while a good balance can be obtained for other criteria. In the light of dynamic change of mining system conditions, a single dispatching criterion is not necessarily reasonable. A complex dispatching criterion has been put forward to seek the complex optimum effect. Taking the advantage of each dispatching criteria, the dispatching result of a complex dispatching criterion dominates over individual dispatching criterion and a comprehensive optimum effect can be reached.

Jan 1, 1995

By Revy T

A testwork program aimed to improve the assaying of underground face samples established that for an Accelerated Cyanidation Extraction using the LeachWELLTM 60X catalyst, followed by an Aqua Regia finish (to quantify non-cyanide-extractable gold): ò Assay results (including the aqua regia finish) from a full set of samples from one underground face can be returned within four hours. ò The reagent dose to make the method simple and 'bullet-proof' is 8 gm LeachWELLTM 60X and 20 gm NaCN per 200 gm sample (30 per cent solids). Pulverising is to 90 per cent passing 75 micron (compared to the standard of 90 per cent passing 150 micron), and cyanidation is completed within one hour - however both might be further optimised. ò The LeachWELLTM/Aqua Regia results demonstrate that Fire Assays suffer from a significant nugget effect which makes the Fire Assay data relatively unreliable. ò Original costs, specific to Bannockburn (viz including an Aqua Regia finish), handling approximately 20 000 to 25 000 per annum is estimated to be $5.40, ie about one-third less than fire assaying. Optimising is further reducing these costs. It is estimated that the LeachWELLT"'/Aqua Regia assay procedure has reduced the probable error limits of any given assay of this ore from the range 0.74x to 1.36x for Fire Assay, to an appreciably smaller range of 0.90x to 1.11x. Thus for a cut-off grade of 1.60 g Au/t, samples with Fire Assays ranging up to 2.18 g Au/t have 25 per cent chance of being below cut-off, while samples with Fire Assays as low as 1.18 g Au/t have an equal chance of actually exceeding cut-off grade. This 1 g Au/t range in uncertainty is generally unacceptable for many practicalgrade/tonnage relationships. By contrast, using LeachWELLTM + Aqua Regia assays, the estimated corresponding limits are 1.44 g Au/t and 1.78 g Auk respectively - a far more tolerable 0.34 g Au/t range of uncertainty. Subsequent data has demonstrated that sufficient accuracy for grade control is achieved by Accelerated Cyanidation alone, further halving the cost of assays.

Jan 1, 1994

Botanical exploration has identified many plant species that naturally accumulate very high concentrations of elements, particularly heavy metals. The enormous potential applications of some of these species to remove metals from soil (phytoextraction) are slowly being realised. The discovery of hyperaccumulation has resulted in considerable research into applications for removing contaminants (phytoremediation) or more importantly, in the context of this discussion, for commercially æfarmingÆ metals (phytomining) from metal-bearing soils. Currently published phytoextraction models estimate Ni yields using Berkheya coddii (a South African nickelophyte) are approximately $US 750 ha-1, excluding the cost of processing. Research indicates that B. coddii will also accumulate significant amounts of Co, resulting in a considerable increase in the value of the crop. Modification of the substrate by acidification, fertilisation and/or chelation indicate fruitful avenues for improving this yield by increasing the extractability of the target element, elevating biomass production rates and/or rendering other contained heavy metals extractable. Plant uptake of more insoluble elements, such as Au, may be induced in many plant species using chemical compounds. These chelates applied to the soil, form soluble complexes with poorly soluble metals forcing them into the soil solution. Plants may then accumulate these metal complexes through transpiration. Research has achieved foliar Au concentrations of 78 mg kg-1 in Iberis intermedia, 56 mg kg-1 in Brassica juncea and 9 mg kg-1 in B. coddii (on a dry weight basis). Therefore in theory B. coddii could phytomine a soil containing modest concentrations of Ni, Co and Au to yield 116 kg Ni, 36 kg Co and over 6 oz Au ha-1 at a current value of $US 4000. Root crops have also shown promise in chelate assisted Au accumulation (eg 6.8 mg kg-1 in Beta vulgaris and 118.5 mg kg-1 in Raphanus sativus). Phytoextraction, in its varied applications, is rapidly coming of age as a technology having considerable potential for application in environmental science and the minerals industry. Phytotechnology employs unique plant biochemical behaviours in the rhizosphere, driven by the sunÆs energy, to seek out and concentrate low levels of heavy metals at very low cost.

Jan 1, 2002

By Smith CB, Boxer GL

The Argyle diamond deposits in the East Kimberley were found as a result of a long term exploration programme, begun in 1972 by the Kalumburu Joint Venture (KJV), to search for diamonds on the Kimberley craton. When CRA Exploration Pty. Limited joined the KJV in 1976, it was re-named the Ashton Joint Venture (AJV). The Kimberley craton was considered prospective by analogy with structural settings of diamondiferous kimberlites in Africa. Diamonds were found during regional drainage sampling in Smoke Creek in late 1979, which led directly to the discovery of the rich Argyle lamproite diatreme (AK1) and associated alluvial deposits of Smoke and Limestone Creeks. The diatreme intruded into quartzites, sandstones and shales of Adelaidean age, and consists predominantly of lapilli ash tuff, with minor coarse and fine ash tuffs, 'non-sandy tuff' and olivine lamproite dykes. Erosion of the diatreme has produced the alluvial deposits of Smoke Creek and Limestone Creek to the north and east. Argyle diamonds are of predominately industrial quality, and the typical diamond is a brown, irregularly shaped, frosted, heavily resorbed form with abundant graphite inclusions, well developed etch channels and prominent hexagonal surface depressions. The pre-feasibility evaluation of AK1 comprised diamond drilling, surface and underground sampling, metallurgical sampling, marketing studies and ore reserve estimation. The proven AK1 ore reserve is 61Mtonne at 6.8ct/tonne. Evaluation of the alluvial deposits gave results of Upper Smoke Creek, 580 000 tonne at 4.6ct/tonne and Limestone Creek, 1.7Mtonne at 3.5ct/tonne. Commercial production from the two alluvial deposits began in 1983 with production for the year reaching 6.2 Mct. Mining of the alluvial deposits at three to five Mct/year is proposed to continue until the start of mining of the AK1 deposit in late 1985 or early 1986 at which time a new processing plant and associated infra- structure will have been constructed adjacent to the AKl deposit in the Limestone Creek area. Production is planned at three Mtonne/year, producing 25Mct/year, which will make Western Australia the largest producer of natural diamond (by weight) in the world.

Jan 1, 1984

By Sale M, Clarke R, Parker M

The Mount Gibson goidmine is located at the southern end of the Retaliation greenstone belt within the Murchison Province, approximately 300 km northeast of Perth, Western Australia. Economic gold mineralization occurs in the upper part of a Tertiary lateritic surface; its ultimate source presumably being Archaean greenstones. Gold is present in the laterite profile over a distance of 4 km in four separate prospects, from north to south, Mount Gibson Well, Midway, Orion and Tobias Find, the two last constituting the Mount Gibson mine. The area has'low relief. A low ridge to the west of the mineralized zone is formed mainly of rubble with minor outcrops of mafic volcanic and intrusive rocks. The terrain slopes gently to the northeast, and the maximum relief, in the vicinity of the mineralization, is about 20 m. The weathering profile comprises, from the surface: (i) soil, windblown sand or colluvium (thin), (ii) laterite; hardcap (including ferricrete and lesser silcrete and calcrete) is overlain by loose nodules and pisolites, and passes downwards into a mottled zone of variably ferruginized clay, (iii) clay; mainly pale brown, cream or yellow in colour, but in places contains relic quartz veins or dark ferruginous zones, (iv) clay with fragments of altered rock; normally grey to pale green in colour but sometimes includes an oxidized iron-rich horizon just above the weathered/fresh rock interface, (v) bedrock. The complete profile varies from 10-100 m thick, averaging 20 m. Zone (ii) is commonly 3-5 m thick. Loose pisolites of this zone fill cavities in the massive ironstone left by former tree roots, and cover irregular depressions which may represent former stream courses. Samples, representing 3 m intervals down the holes, have been collected from cuttings taken from rotary drill-holes along each of four traverses; one traverse at each-of the northern prospects and two traverses in the Orion part of the mine area. These samples have been submitted for X-ray diffraction analysis and for chemical analysis for a wide range of elements. Bedrock lithologies have been determined from microscope studies of bottom-of-hole chips.

Jan 1, 1988

By A L. Hemphill

Conventional models, based on pulp kinetics, were regressed directly to steady-state data from a nine-stage platinum flotation plant, to provide a basis for optimisation studies. æFroth factorsÆ were used to characterise the froth in each stage and a ædepressant factorÆ was used for gangue minerals, to account for addition of depressant to the cleaners. These factors improved the fit, but differences between plant and model of up to 50 per cent were still present in a number of streams in the cleaning circuit. A new model was developed, which incorporated two rate constants for the recycle of particles from the froth. These parameters provided a workable approximation of the distribution of probabilities of particles being recycled in plant cells, depending upon where particles appeared at the surface. The feed streams were characterised by two pulp kinetic parameters and the above recycle parameters, for all minerals and all stages. The fit to the data was improved very significantly, when compared to the use of conventional pulp-based models with several pulp rate parameters. This relatively simple model did not require additional data from batch tests and it was suitable for plant simulations.Simulations of various strategies for plant operation were tested, allowing the simulator to vary the concentrate flows from all or some of the stages, in order to maximise platinum recovery. The simulator was imbedded in an automatic optimiser and constraints on chrome content or mass flow of final concentrate were used. The simulations of the existing circuits demonstrated that an improvement in recovery, (for a constant concentrate mass) of about 0.5 per cent was possible, by optimising all concentrate flows. A significant reduction in the flow of scavenger concentrate and an increase in the recycle flows in the cleaners were noted. The results of simulations of various changes to the circuit configuration are also discussed. An increase in recovery of 1.4 per cent (for constant concentrate mass) has been achieved in recent simulations, through circuit changes. However, it was necessary to keep the mass pull of concentrate, per unit area, within typical limits.

Jan 1, 2005

By J J. Jiang, M Reimnitz

The Kalgoorlie Consolidated Gold Mines (KCGM) Fimiston Open Pit is mining through a dense network of disused level workings and open stopes of greatly varying sizes. The condition of the stope crowns and remnants is often unknown, and ultimately these structures become a significant proportion of the production floor and pit walls. Significant seismic events and occasionally rock bursts often occur in the underground workings. Seismicity is becoming an issue to the viability of the KCGM Fimiston operations. KCGM needs to take a leading role in understanding mine-induced and regional seismicity. The seismic system trials were initiated in the Fimiston Open Pit in 1996 using the ISS system. The initial system included two triaxial geophone arrays and two triaxial accelerometer arrays installed into two boreholes. The main objectives of the trial were to investigate the suitability of the two types of sensors and the detectability of seismicity in the pit. The trial was regarded as a success and the system has been expanded and improved. The current system consists of 12 triaxial geophone and accelerometer arrays and nine uniaxial geophones. The 12 triaxial arrays are installed in nine boreholes drilled from the pit and the nine uniaxial geophones are installed from underground level workings about 660 m below the original surface elevation. The system is designed to provide seismicity coverage for the entire pit area that is about 2.7 kilometres long, 1.5 kilometres wide and about 300 metres deep. The accuracy of the source location has always been a challenge due to the existence of the dense concentration of stopes and lack of the sensors. It is uncommon to have direct ray path for the seismic waves to propagate to the sensors across the strike, hence sensors have been located along strike and under the old workings. This paper discusses some of the challenges associated with seismic monitoring near mine remnants and seismically active faults. The results of the seismic monitoring are presented, and contrasted to results of three dimensional boundary element numerical modelling. The impact of the seismicity upon pit slope stability is briefly discussed.

Jan 1, 2003

By D A. Shushkov

Analcime-bearing rocks are one of the varieties of raw zeolite material, although the characteristics of analcime are not as good as other kinds of zeolites such as clinoptilolite or mordenite. However, some fields of application of analcime-bearing rock have been found, eg a method of water purification has been developed and patented. Analcime-bearing rocks are rather widespread in the Timan region of the Komi Republic of the Russian Federation. They are represented by Upper Permian siltstones and mudstones. The analcime-bearing rocks have the following mineral composition (per cent): clay minerals (50 - 70 per cent), quartz (ten to 30 per cent), analcime (one to 30 per cent), feldspars (two to ten per cent), carbonates (two to five per cent) and pyroclastic material. Analcime aggregates encrust cavities and infill microfissures in rock. Both microcrystalline (crystals <0.01 mm in size) and medium-crystalline (~0.5 mm) aggregates are observed. The mechanical, gravitational, magnetic, sorption and ion-exchange properties of analcime-bearing rocks have been studied. They are characterised by a high index of grindability and a low index of abradability and crush resistance, low values of the benzol-based pore volume and toluene-based specific surface area, and a medium porosity value. The magnesium based ion-exchange capacity (IEC) ranges from 28.7 to 196 mg+equiv/100 g. Analcime-bearing rocks of the Timan region are characterised by low beneficiation potential. Preliminary ultrasonic treatment of the analcime- bearing rocks is efficient for the destruction of intergrowths along their boundaries and production of free analcime aggregates. The proposed beneficiation procedure yields 95 per cent analcime concentrate. In the Komi Republic, analcime-bearing rocks are utilised in agriculture. They are also used for domestic needs (eg as biotoilets for domestic animals) and the solution of ecological problems (regeneration of soil polluted by oil). We have completed experiments on the purification of potable and industrial water using analcime-bearing rock subjected to thermal treatment. The experiments demonstrated that it was possible to decrease water colouration by 20 per cent, turbidity by 79 per cent, permanganate oxidation by 30 per cent and total iron content by 70 per cent.

Jan 1, 2008

By Dawes S

Efficient roof control in colliery operation is a factor that must be maintained for· the maximum reQovery and economic extraction of the mineral. It must be carefully considered in each of the following stages in the life of a mine.1. The primary development and the first working of the mine.2. The maintenance of access roads, intake and return airways.3. The extraction of pillars.FACTORS IN ROOF CONTROLThe main objects of roof control are: The support of the roof in working places, the relief of roof pressure from the workings during pillar extraction, and in some instances the prevention of subsidence of the surface.The support of the roof during development operations and first working, and the subsequent maintenance of these is influenced by:-1. The depth of the seam below the surface.2. The thickness of the seam.3. The dip of the seam.4. The condition of the roof and floor.5. The method of extraction-mechanical or by hand.6. The method of. working-Iongwall or bord and pillar.7. The availability of material.As a general rule, the pressure exerted by the superincumbent strata may be taken as one pound per square inch for each foot of cover, and the crushing strength of coal as 1,500 to 2,000 lb. per square inch; this, of course, varies, but from these figures it may be assumed that seams below 2,000ft. in depth are liable to crush. If any part of the seam is removed, added weight is thrown onto the remaining coal. It is therefpre essential that for effective working this extra weight must be removed or supported as soon as possible.The outstanding principle in the support of mining excavations is to ensure that the supports are set as soon as possible after the natural support has been removed.Where a system of bord and pillar is applied, the early setting of supports reduces the extent to which the roof bends, thereby reducing the liability to ·tension fractures at the centre of the road.

Jan 1, 1947

In the 1890s, The Goldfields Water Supply Department in Western Australia commissioned the diamond drilling of two deep artesian exploratory water bores near Coolgardie (Blatchford, 1899). At the time, in spite of strongly-worded advice to the contrary from the Government geologists, there was a wide popular belief that you could obtain artesian groundwater virtually anywhere providing you drilled deep enough. The first Coolgardie hole was drilled to over 900 m, despite the fact that the bore penetrated granite from a depth of 5 m below the surface. Drilling proceeded for more than 16 months, between July 1896 and November 1897. A small quantity of salt water was struck at a depth of 36 m. The second bore was mercifully terminated at a depth of 50 m, also in granite, with again only a small inflow of salt water. Such were the desperate lengths that were pursued to obtain a reliable water supply for the influx of people in the Eastern Goldfields gold rush. The WA Goldfields continued to rely on huge condensers for water supply until that great engineering feat, the construction of a water pipeline from Mundaring Weir to Coolgardie, was completed in 1903. An enormous network of rail tracks penetrated outwards from Kalgoorlie and Coolgardie for the collection of timber, principally to feed the condensers. Remnants of the rail network are still visible today. The acute shortage of fresh water led to great deprivation, disease and death among the early miners. Today, the water problems faced by the mining industry are thankfully not as severe, but they are nevertheless no less challenging. The southern part of the WA Goldfields is a notoriously difficult area for finding potable quality groundwater. Even as recently as the 1990s, there were occasional serious efforts to find potable groundwater in that area, some very imaginative and innovative. One mining company, working south of Southern Cross, spent many months experimenting with recovery of potable quality water from a shallow freshwater lens overlying hypersaline water in a low permeability weathered granite aquifer. They used very low capacity pumps manufactured from car windscreen washer pumps powered by solar panels.

Jan 1, 2000

By H R. Phillips

Fresh air intake in an underground coal mine is of critical importance to the occupational health and safety of workers due to the dynamic, confined and dangerous nature of the mining environment. In most coal mines in South Africa, belt road is not separated from the fresh air intake to the sections. Intake air to coal production faces generally contains dust generated by conveyor belts, transfer points and travelling roads. Long roadways (up to a few kilometres in some of the existing mines) to the coal sections are involved in the ventilation circuits of individual sections. This leads to a progressive increase in the dust levels in the intake fresh air and thus high dust levels in the air delivered to the working sections downstream. In the process of reducing worker exposure to dust in the face area, attention to the monitoring of fresh intake air is generally given a low priority. This paper summarises the intake dust levels measured in selected South African underground coal mining sections over the past five years and its use as a parameter in assessing worker exposure to dust through a newly developed model called the Dust Exposure Level Index (DELI). Based on the measured data of the South African coal mines, the average section intake dust level was 0.78 mg/m3. Approximately 60 per cent of the samples collected for the study have exceeded the 0.5 mg/m3. The intake dust level measured at a longwall face was much higher than in the CM headings. The measured section intake dust levels furnish the South African coal mining industry with the latest depiction of the level of face area dust contamination due to the intake air. Further, this paper demonstrates the use of intake dust level as a parameter in estimating exposure levels of workers, as it clearly plays as a corner stone in effective control of exposure to face area respirable dust. It must be borne in mind that intake dust is a base level to the mine worker exposed to dust without carrying out any additional work in the section. If the intake air is seriously contaminated by airborne respirable dust; then the exposure of the workforce in the section will constantly be elevated by that background dust level.

Jan 1, 2005

By C G. Ferguson, S L. Makin

The Cawse nickel project is a new generation pressure acid leach (PAL), solvent extraction and electrowinning process used to produce LME grade nickel metal and cobalt sulphide concentrate containing 40 per cent cobalt by treating nickel and cobalt oxide ores from a lateritic ultramafic profile. The project is located 50 kilometres NW of Kalgoorlie in Western Australia. The commercial success of the Cawse nickel operation is dependent in part upon the ability of the principal ore type to be beneficiated. A simple process of rejecting lower grade material on the basis of particle size achieves this. The process is called upgrading and allows subgrade material to become commercial ore grade and also allows higher ore grades to be processed in the early parts of the operation. The success of the upgrade process is dependent upon the physical and chemical properties of the ore being processed and the nickel grade of the ore. The upgrade ore type represents >80 per cent of the Cawse nickel resource and is hosted mostly within a limonite unit, represented by goethitic clays in the upper saprolite zone. In the upgrade processing circuit the ore is crushed and slurried. It is then passed over a series of screens that remove all material over 0.5 mm. This ensures that coarse material is removed from the main ore stream. Upgrade is reliant on selectively rejecting lower grade or barren material, which can include silica, magnesite and relict saprock. To predict leach feedgrade, the grade control estimates the head grade and upgrade percentage for all ore delivered to the run of mine ore pad. To do this reverse circulation (RC) drill holes are logged for ore type, assayed for a range of elements and undergo simulated upgrade testwork. The percentage and grade of material rejected through the upgrade circuit has the most influence on upgrade per cent. This reject is determined by the physical properties of the rock mass and drum scrubber parameters. The correlation between nickel upgrade and reject can be established to predict upgrade from the known mass reject. Parcels of upgrade ore milled to date closely follow projected correlation trends. Upgrade circuit results have been excellent, with average nickel grades upgrading by 34 per cent, cobalt by 23 per cent, with 55 per cent mass passing.

Jan 1, 2000

The aquifer at Phosphate Hill, north-west Queensland, presents special management challenges. The site contains a phosphate rock mine, fertilizer plant and camp. All of the site water is provided by the Duchess Embayment Aquifer, located within the Burke River Outlier of the Georgina Basin. The Duchess Embayment is a triangular shaped, Cambrian marine sedimentary basin consisting of siliceous and calcareous phosphorites, shales, siltstones and sandstones. It is fault-bounded on all sides and has a regional dip of 5¦E. The aquifer is largely confined to the siliceous facies of the Beetle Creek Formation. It has a very flat potentiometric surface and an average thickness of 45 m with confining beds above and below. The siliceous facies derives from weathering of the calcareous phase and is highly transmissive while the calcareous zone itself contains less water and is dependent upon interconnecting fractures for its permeability. Recharge appears to occur where the siliceous facies crops out around the edges of, and within, the Embayment. Aquifer management issues revolve around two points: successful dewatering of the orebody for mining; and efficient management to maintain water quality and ensure supply over the predicted 30-year life of the project. Current operations are extracting water at levels much below that originally estimated. This is having a detrimental effect on mine dewatering with the SWL lowering at a rate slower than anticipated. As a result, ore parcels are now beginning to intersect the water table, providing a major constraint to the mining plan. The aquifer drawdown cone is extensive and very flat so to keep the water table at an appropriate level for mining the management plan will include more aggressive dewatering to lower the level of the entire aquifer. The additional water will either be disposed of downstream or stored in a holding dam or via the use of artificial storage and recharge. Water resource management will also need to maintain the current quality. About 37 per cent of the water currently extracted forms the low chloride supply required by the ammonia plant and the power station. This ratio is expected to remain applicable under any future operating conditions but the low chloride resource is restricted so will require prudent management to enable the supply to continue. At current usage rates aquifer recharge versus use will need careful monitoring to ensure that the resource remains viable in the long-term.

Jan 1, 2003

By B Ilich, M Brad, R Brandli, E MarT

The Wirtgen surface miner was investigated for application to the Cloud Break iron ore deposit since it can selectively mine thin seams and there is no requirement for drilling and blasting or primary crushing as the surface miner cuts and breaks the rock to secondary cone crusher feed size. A physical metallurgical study of the surface miner run-of-mine (ROM) from Cloud Break was conducted to better understand the product size distribution, amount of -31.5 +6.3 mm lump produced, lump stability and physical metallurgical quality compared to conventional drill and blast techniques. Since the Cloud Break mine was not in production, the lump product from crushing and screening of surface miner ROM was compared with product simulations using traditional premining physical metallurgical processing techniques including drop testing of PQ diamond core. Drop test conditioning is widely used in the iron ore industry to simulate blasting, crushing and screening at the mine and subsequent product degradation during handling from mine to port to predict both the amount of lump product on ship and the size distribution of the fines and lump. The size distributions of surface miner ROM for hard, medium and friable ore types show very tight size distributions with very little above 100 - 150 mm, about 60 per cent of the distribution between 75 and 6.3 mm, with less below 1 mm compared to conventional drill and blast ROM. Crushing, screening and drop test conditioning of the surface miner ROM produced more lump in equivalent ore types compared to drill and blast techniques or their breakage simulation using PQ diamond core. The higher lump yield from the surface miner was also confirmed by pilot scale crushing of some 7400 t of surface miner ROM from the Cloud Break trial pit. The lump produced from surface miner ROM was quite stable and showed similar low degradation characteristics to lump simulated from drill and blasting techniques. Product size distribution simulations using conditioned crushed lump from 1 m diameter Bauer hole samples were found to compare well with processed PQ core samples and surface miner product from trial mining at Cloud Break. The Bauer sample method is therefore considered to be a very useful close simulation to actual lump production, although the lump yield is lower.

Jan 1, 2007

The Ranger Project Area covers 78.6sq km and is approximately six km east-west by 14km north-south (Fig. 1). The area is known to con- tain two uranium orebodies, Ranger No. 1 and No. 3, two uranium occurrences, Anomalies 4 and 5 and three significant radon anomalies, Ranger 18, 19 and 48. The Ranger No. 1 and No. 3 Orebodies con- tain approximately 120 000 tonnes of U308 in highly chloritised metasediments of the Lower Proterozoic Cahill Formation. The orebodies occur as suboutcrop and exhibit significant radiometric anomalies. These occur within approximately 400m vertically of the projected unconformity with the sandstone of the Middle Proterozoic Kombolgie Formation. Oxidised and reduced uranium minerals occur in quartz-rich schists containing variable muscovite, chlorite and sericite and in the chloritic matrix of chert breccia. Secondary uranium minerals namely saleeite, metatorbernite and gummite associated with earthy iron oxides occur from the surface to 50m depth in No. 1 Orebody and up to 80m depth in No. 3 Orebody. Below the oxidised zone, uranium occurs as pitchblende finely disseminated throughout sequences of quartz muscovite chlorite schist up to 150m thick and as open space fillings in zones of structural deformation. The uranium-bearing schists directly overlie up to 300m of dolomite and magnesite containing minor interbedded schist units and are themselves overlain by approximately 2 000m of more micaceous schists of the upper part of the Cahill Formation. Regional metamorphism has occurred to at least amphibolite facies and has been followed by retrograde metamorphism to a greenschist facies equivalent. The metasediments are intruded by granitic pegmatites and dolerites. Uranium is associated strongly with chlorite of which there are at least five textural varieties. Not all of the various generations of chlorite have associated uranium. Graphitic schist horizons occur which are often less than one metre thick and discontinuous. These are best developed at the base of the quartz chlorite schist sequence where uranium grades up to 1 000ppm U may extend for several hundred metres beyond the main ore zone. Uran- ium is not common in the carbonate sequences and only occurs in chlorite segregations in disrupted zones. The unoxidised ore contains sulphides, mostly pyrite, which are low in abundance (

Jan 1, 1984

By J Rutter, M NoppT

Ridgeway Gold Mine is part of Cadia Valley Operations which is wholly owned and managed by Newcrest Mining Limited. Ridgeway is an underground sublevel cave (SLC) operation mining approximately 5.6 Mt of gold-copper ore per annum. Ore tracking and accounting systems monitor gold and copper grades, copper mineralogy, host lithology and ore sizing. These factors have material impacts on concentrator throughput and metal recovery. Drawpoint recovered grades are routinely reconciled back to design data to monitor and assess draw performance. Grade control at Ridgeway has been based upon sampling the ore from SLC drawpoint bunds using 15 kg grab samples combined to form 30 kg samples prior to assay. This large sample size was selected in an attempt to representatively sample the coarsely-sized broken ore, the samples collected having a nominal diameter of about 130 mm. The assay results for gold and copper from these samples are extremely variable and the usefulness of the results has been questioned by the mine, particularly given the onerous and OH&S hazardous activity of handling large sample masses. Granulometry testwork at Ridgeway confirmed the extent of the relationship of decreasing sample grade variability with particle size. Based on this relationship, drawpoint sampling tests were carried out using 9 kg samples sieved to -45 mm at the drawpoint site. The gold and copper results from the finer 9 kg samples were significantly less variable than those from the normal, coarser 30 kg samples. The variance reduction was similar to that predicted from broken ore sample theory for the selected sample mass and nominal particle size. This relationship therefore allowed Ridgeway to assess the sample mass and particle size required to achieve a particular precision. In addition to improving the precision, the reliability of assay results from smaller samples from the fine particle fraction for drawpoint control depends on the accuracy of the results. The assumption is that the grade of the finer ore is similar to that expected from the lot, or that if a bias exists, then the extent of this bias is acceptable and consistent enough to be accounted for in the grade control system. The distribution of grade with particle size and the reconciliation of drawpoint sample and mine results is currently under investigation.

Jan 1, 2005

By Hagarty KA

At the northern margin of New Guinea there is a fundamental problem related to the relative timing of arc- continent collision(s) represented in the Papuan Fold-Belt (PFB) and backarc spreading in the Caroline basins. New fission track dates from basement and sediments within major anticlines of the PFB indicate that uplift occurred at 4.0 ¦ 0.5 Ma. The timing is consistent with the fold and thrust shortening (more than 100 km) of the PFB which occurred in the Late Miocene/Early Pliocene as evidenced by the deformed Miocene carbonates and Pliocene molasse deposits. However, in northern Papua New Guinea, earlier work has suggested that ophiolite emplacement and associated metamorphism occurred in the late Oligocene/early Miocene.The evidence suggests that the late Miocene/early Pliocene deformation was of considerably greater magnitude than the earlier obduction event. The Caroline Plate lying north of New Guinea has been interpreted as a backarc basin which commenced spreading in the Eocene and ceased spreading in the Late Oligocene (about 28 Ma). The remnant arc presently docked to New Guinea was active in the Eocene and Oligocene and is believed to be the arc related to the formation of the Caroline Plate. The cessation of spreading of the Caroline Plate has previously been interpreted to be due to arc continent collision during the Late Oligocene/Early Miocene, but such a collision is now known not to have occurred. In the Cretaceous and Palaeocene subduction was to the south beneath New Guinea, but "flipped" in the Eocene to give a north-dipping subduction zone far to the north of New Guinea, driving Caroline Plate backarc extension. This backarc basin reached a limiting size at about 28 Ma resulting in the end of subduction. With the continued northward movement of Australia, Mesozoic oceanic crust was obducted southward onto the New Guinea margin in the Late Oligocene/Early Miocene. In the Miocene south-dipping subduction beneath New Guinea commenced until the now extinct Caroline Plate arc collided with New Guinea in the Early Pliocene, causing extensive regional compression and sinistral strike-slip faulting. This new tectonic framework suggests that backarc basins reach a limiting ridge-trench distance of 1000-1500 knm, beyond which "subduction-pull" cannot drive further spreading so subduction and spreading are terminated.

Jan 1, 1987

The Malvern Hills coal project is now in an advanced stage of development with the stripping of overburden from above former underground workings in which a 3 m seam of subbituminous C coal with very low ash and sulphur was previously mined. The main seam dips at 45¦ to the south-east, along with other adjacent potentially recoverable thinner seams, and the strike extent of past coal mining in the Malvern Hills approaches 10 km with most extraction by underground methods. Overburden and interburden is predominantly weak mudstone, with only very minor sandstone, whilst a 1 - 2 m thick veneer of Last Glaciation loess covers the area. The coal product from the present mine site will supply the local industrial market, with production planned to peak at about 20 000 tpa. Settling pond design has been based on several different approaches, including an estimate of likely inflows from the contributing catchment of about 5 ha under predicted storms having a ten year recurrence interval. A pond of about 1000 m2 in area has been constructed with its invert in mudstone bedrock, and soil slopes have been lined with lime-stabilised clayey loess silt incorporating a gravel armour layer. The pond outflow takes place via a 2 m high 1.8 m diameter concrete dewatering tower, with a skimmer arrangement to lower the pond by approximately 500 mm between storm events. The overflow water discharges to the main stream above a natural wetland, which will act as a secondary polishing pond, and this is to be enhanced in conjunction with the neighbouring forestry company. Discharge consents specify turbidity and pH as the controlling criteria, and both will be monitored at the settling pond discharge point as well as at the entry to the main stream. Investigations to date indicate that the pH of water exiting from previous underground workings is in the range 6.8 to 7.2, and that acidic drainage will not be an issue with this mine site. No dispersive clays have been identified in assessing the overburden and batter materials, although erosion during storm events is expected and all drains will ultimately be routed through the main settling pond to ensure adequate control of runoff-generated sediment. Reestablishment of vegetation on cut and fill batters remains a priority, and ongoing monitoring will ensure satisfactory long-term performance of the pond system.

Jan 1, 2003