Search Documents

The five main ways in which coal is used are1. for electricity generation,2. as a lump fuel for domestic and industrial purposes,3. for gas making,4. for making carbon for chemical and metallurgicalindustry, and5. as a source of coal tar and coal-tar chemicals.In the past about 80 per cent of Victoria's brown coal has been used for power generation and the rest for making a lump fuel, in the form of briquettes. Some of these briquettes have been used for gas making, and a carbon industry based on them has recently been commenced, and could expand rapidly. However the enormous increases expected in power requirements, approximately a doubling every 10 years, will ensure that even if the manufacture of carbon increases dramatically and a large amount of coal is used for other purposes, the brown coal industry will always be dominated by power generation. As new fuels such as natural gas, oil, and nuclear fuel become available at competitive prices these will take their share of the power generation load.Natural gas or oil, or possibly pumped storage hydro, will almost certainly be used for all future peak-load stations, comprising around 40 per cent of the total installed generating capacity, but new base load stations will continue to be designed for brown coal operation until about the mid-1980s, by which time nuclear power stations will become competitive. By this time Victoria will have installed about 6000 MW capacity of brown coal stations, burning around 50 million tons/year of coal, compared with 18 million tons/year at present, with perhaps 15 million tons/year being used for other purposes.This consumption of 65 million tons/year should then remain fairly stable over the life of the power stations installed, well into the 21st century. Brown coal therefore appears to have an assured place in providinga major part of Victoria's fuel requirements, despite the diversification in fuel usage pattern which appears certain to occur with the recent developments in new fuel resources.

Jan 1, 1971

A static equilibrium mathematical model of the lead,flotation circuit at The Zinc Corporation Limited is developed to permit the feasibility of applying computer control techniques to be studied. The model is based on the relationship between the cumulative recovery of a mineral species R from n flotation cells or assemblies in a unit given asR = M(1-(1 - p)?)where M is the maximum possible: recovery and p is the probability of flotation for the particular mineral.The verification of this relationship by data from pilot plant cell-to-cell and plant open flow flotation units shows that these systems behave as if the probability p is constant for each cell or assembly of anyone unit. It is also concluded from this study that mineral recovery is independent of the head assay. Two consequences of this probability relationship, are demonstrated.1. The total recovery from an open flow unit is enhanced by increasing the number of flotation assemblies but maintaining a constant flotation time and total air addition.2. The rougher and scavenger capacity required to recover a bulk lead-zinc concentrate is approximately half that used for selective flotation.The static equilibrium model is derived by solving a series, of mass balance equations which incorporate the recovery-probability relationship for each type of flotation unit. It is used to calculate the complete metallurgy of the circuit for any arrangement of cells knowing only the plant feed assays and the probabilities of flotation for the three mineral types, galena, marmatite, and gangue for rougher, scavenger, and cleaner flotation. The high degree of prediction obtained is demonstrated by comparison with plant results.Pulp flowrate is the principal variable affecting the accuracy of the present model. Work is in progress to establish this effect and permit the derivation of a dynamic model. Even so, the static model is valuable in its present form for the analysis and design of circuits, and optimisation studies using computer services have been conducted.

Jan 1, 1964

By W A. Ducker, R-H Yoon, M Mao

Froth flotation relies on an attractive force between a particle and a bubble to selectively extract particles from a mixture. The range and magnitude of this attractive force is an area of ongoing investigation. We describe measurements of the effect of dissolved gases on the long-range attractive surface forces between a silica plate and a glass sphere in aqueous octadecyltrimethylammonium chloride (C18TACl) solutions. The measurements follow on the work of Sakamoto et al (2002) who have concluded that long-range interactions occur in air-equilibrated surfactant solutions but not in degassed solutions. In the present work, we have measured attractive forces that have a much larger in magnitude than the expected van der Waals force in solutions that were degassed by the same procedure used by Sakamoto et al (2002). We find that the forces are most attractive at 5 + 10-6 M C18TACl, which is close to the charge compensation point (ccp) of the glass sphere. At this concentration, the attractive force decays exponentially with distance and the decay length is 34 in air-saturated solution, and 38 nm in degassed solutions. In both air-equilibrated and degassed solution, the force is quite sensitive to the concentration of surfactant. The force is also sensitive to the pH. We note that our degassing procedure increases the pH of the solution from about 5.7 to 6.6 (presumably by removal of dissolved CO2). We find that the forces measured in degassed solution can be reproduced by using a solution that was equilibrated in air, then the pH was increased by addition of NaOH. Therefore, we conclude that the main effect of degassing in our experiments is to raise the pH. We show that the range and magnitude of the attractive force decreases in the presence of salt (NaCl). This suggests an electrostatic mechanism for the attraction. Possible mechanisms for the attractive force are examined, including the mechanisms that rely on a patchy coverage of surfactant on the silica surface.

Jan 1, 2005

Hindered settling devices (teeter bed, up-current) are widely used in the mining industry to classify mineral streams into various size fractions and to separate minerals. In the coal industry these devices are extremely successful in separating coal of low densities from shale and pyrite. In the beach sand and in the iron ore industry, this equipment is used to beneficiate minerals either as stand alone units or lately as classifiers in spiral circuits. The challenge that these applications give to metallurgists is to be able to simulate the equipment by using mathematical models with or without the aid of small bench scale tests. In any green fields and many brownfields projects, limited amounts of samples make pilot plant tests difficult. Although there are mathematical approaches available, the accuracy is of such a nature that it can not be used for design purposes. This refers to the use of StrokeÆs law, NewtonÆs and the so called intermediate law as is widely published. Research done by the author indicates that in most applications none of the above laws apply. Using the original data as published by Perry and Chilton (Chemical Engineering Handbook, pp 5-64, 1973) and re-plotting it indicates that the settling velocity of particles is a function of (particle diameter)n where n varies between 1.2 and 1.6 which is different from that predicted by the intermediate law that suggests that n = 1.18. Another way of describing the relationship between the drag force and the Reynolds number is done by iterating between settling velocity and the Reynolds number. This needs more computer time but the accuracy obtained from this does add value to simulations. A small but simple test is suggested to determine the settling velocity of particles to enable metallurgists to determine variations from the theoretical calculations to determine the influence of viscosity and particle shape.

Jan 1, 2008

By K Gottliebsen

Many hydrometallurgical and chemical processes produce large amounts of acid waste. This is illustrated by the copper S/X-E/W process used at the Mount Gordon mine in North Queensland, which produces a tank house bleed in the form of a concentrated sulphuric acid stream. The bleed stream is necessary to remove impurities such as iron that decrease the efficiency of the electrowinning process. However, this bleed stream comes at a high financial cost due to large quantities of make up sulphuric acid required as well as the need to neutralise the bleed stream prior to disposal. The present work describes the approach taken to develop a solvent extraction process for the selective recovery of sulphuric acid from copper tank house bleed liquor. The proposed extraction system was investigated using a branched long chain aliphatic tertiary amine tris(2-ethylhexyl)amine (TEHA) as the extractant, Shellsol 2046 as the diluent and octanol as a modifier. Computer simulations were used to determine the optimum operating conditions for the recovery process. It was found that the process was able to selectively recover up to 90 per cent of the sulphuric acid from the copper electrolyte. This is recovered as a pure aqueous acid stream at a concentration of up to 130 g/L that can be recycled back into the tank house circuit or other areas of the process. The raffinate acid concentration was reduced from 180 g/L to as low as 18 g/L, significantly reducing waste disposal and the environmental impact of the copper process. The equilibrium data and simulation results were also compared with an alternative extractant, CYANEXR 923. Finally the efficiency of the TEHA extraction system was investigated on a pilot-scale in a pulsed perforated plate column. This information was used in order to scale the process for an industrial application. The same methodology applied in this case study could be used to provide innovative and cost-effective solutions to waste streams encountered in other hydrometallurgical or chemical processes.

Jan 1, 2000

By Wang S

Fragmentation is the basic concern in rock blasting and serves as the main measure of blasting effectiveness. A good estimation of fragment size distribution before blasting is very useful and helpful for selecting the proper equipment for loading, transporting, and crushing rocks. A fragmentation model is presented that is based on a hole by hole analysis and correlates the rock properties, burden, spacing, and explosive column with the blasted volume and fragmentation obtained. The drill bench is divided into three separate areas corresponding to different breakage mechanisms. They are: 1) a main breakage or casting area, 2) a secondary breaking area, and 3) an overbreak area. The boundary of the main breakage area is found from the magnitudes of strains in the stress waves and the tensile strength of the rock. The secondary breaking area is determined by stress wave reflections at discontinuities and interactions between colliding stress waves from adjacent boreholes. The overbreak area is that beyond the main breakage area "where the reflected tensile waves no longer break the rock. This analysis is applied on a hole by hole basis to locate the breaking points. Then connecting the breaking points on each side of a borehole forms the triangle of the main breakage area. This is repeated for the next hole and the remaining area is the secondary breakage area. After the angle and area for each hole are found the Kuz-Ram model is used to predict the fragmentation size distribution for each hole. Then summing up the size distribution for each hole gives the total fragmentation distribution. This analysis has been successfully applied to field tests performed in a limestone bench where the entire spoil pile was screened to determine the fragment size distribution. Knowing the explosive and rock properties the breaking area was found for each hole, the strain and stress values, the breaking point, and angle of breakage (104 degrees). The fragmentation size distribution was then calculated and compared to the field data with excellent agreement.

Jan 1, 1990

IN the ordinary course of treatment of pyritic ores or concentrates, whether for the recovery of copper, tin, gold, or any other contents, the removal of some or all of the sulphur contents is almost invariably the first step. The practice is to roast the ore in a current of air so that the sulphur is oxidized to sulphur dioxide and the metal constituents are changed from sulphides to oxides. If it is feasible to produce sulphuric acid in conjunction with the treatment process, the sulphur can be usefully employed; but if this is not possible, then the sulphur contents not only become a dead loss, but tend to be a nuisance as well in the surrounding neighbourhood.The problem of avoiding this disadvantage and making economic use of the sulphur has had attention from the early days of scientific metallurgy, and in the past few years very special efforts have been made both to drive part of the sulphur as such from pyrites and recover it, and to replace the sulphur in the S02 by carbon, with the production of carbon dioxide and precipitation of the sulphur. So far economic success does not appear to have crowned these efforts. In the case of pyrites, of the two atoms of sulphur which go to make the FeS2, the first is very lightly held. Dissociation is effected by heat alone at a temperature which is given by Peters as 5160 C.Now, sulphur and hydrogen will unite or burn together, with the production of sulphuretted hydrogen. It occurred to the writer that, if hydrogen was passed over pyrites heated to the above temperature, this first atom would be snatched away by the hydrogen, and could possibly be economically recovered.The first step consisted of some simple laboratory experiments. From these it was found that this first atom was readily removed, and that the extraction was practically complete. Further, the reaction took place at a low temperature, a dull red heat being sufficient.

Jan 1, 1923

By J D. Bryant

The Ernest Henry Copper-Gold mine, the third largest copper deposit in Australia, is located 35 km NE of Cloncurry in the Mt Isa-Cloncurry mineral district of NW Queensland. Ernest Henry Mining Pty Ltd (EHM), a wholly owned subsidiary of Xstrata Copper Australia, operates the mine. Approximately 10.2 Mt of ore grading 1.12 per cent Cu and 0.56 g/t Au were processed during the 2001 - 2002 financial year. The Ernest Henry plant produces copper-gold concentrate, which is trucked to Mt Isa for smelting. The Ernest Henry deposit is a member of the diverse Fe-oxide-(Cu-Au) class of deposits. The orebody is hosted within a sequence of moderately SSE-dipping, intensely altered Paleoproterozoic intermediate metavolcanic and metasedimentary rocks. Copper (chalcopyrite)-gold mineralisation occurs mainly within the magnetite-biotite-calcite ¦ pyrite matrix of a 250 m + 300 m pipe-like breccia body. The breccia pipe comprises two anastomosing lenses separated by a lower grade, clast-supported breccia zone. Several anastomosing brittle faults affect the macro-geometry of the orebody. Copper grade is proportional to the degree of matrix support in the breccia. The July 2002 Ore Reserve was 88.7 Mt @ 1.0 per cent Cu and 0.5 g/t Au. Resource estimation at Ernest Henry is based on a large database of predominantly diamond drill hole information. Drill spacing is typically 50 m + 50 m. Ordinary kriging is used to estimate Cu, Au, SG and IST (oxidation indicator) into model parent cells of dimension 20 m E + 20 m N + 16 m RL. The block model is comprised of five geological-geochemical domains represented by 3D wireframe surfaces. Zonal control, where samples and blocks are assigned the same domain number, is used to control grade estimation. The geometry of the domains defines hard grade boundaries across strike and soft boundaries along strike A Deep Drilling Project was initiated in August 2002 with the initial aim of testing the down-plunge extension of the orebody to a vertical depth of over 1 km. The encouraging intersections to date have resulted in a significant increase in the mineÆs open pit Mineral Resource together with the identification of an underground Mineral Resource. Added benefits of the project included an expanded and updated geological legend and an improved understanding of the host geological sequence.

Jan 1, 2003

By K Baskaran, T McKenzie

Sustainability is becoming increasingly important in the mining and mineral processing industries and must incorporate the associated waste products. Acid mine drainage (AMD) is one such by-product and is one of the most serious environmental problems facing the minerals industry today. The oxidation of sulphidic mine wastes often continues for a substantial period of time after mine closure, resulting in difficult and costly remediation and rehabilitation works. Mining companies are often reluctant to spend increasing amounts of money on waste treatment when the mine life is limited or even finished. Hence a simple, low maintenance and low-cost method of treating AMD is required. Whilst this paper does not address the issue of AMD, it does propose methods for removal of individual species from AMD with potential benefits, including raising AMD pH. A novel concept of using biosolids as a biological adsorbent, or æbiosorbentÆ, of metals from AMD is being investigated at a laboratory/pilot scale level. Biosolids are a by-product resulting from the biological treatment of wastewater, and have been previously shown to adsorb metals from aqueous solutions. This could lead to an environmentally sustainable or ægreenÆ method for treating both AMD discharges and disposing/reusing the biosolids. The result of a laboratory-scale study of the biosorption of Zn(II) is presented in this paper. Physical parameters including reaction kinetics, mixing speed and solution pH were investigated. Solution pH also rose an average of 2 pH units over the 24 hour equilibrium time û a valuable side effect when treating acid mine drainage. The outcome of the study highlights the usefulness of biosolids as a biosorbent for the removal/recovery of metal ions from acid mine drainage. A simple, low-cost treatment technology requiring low maintenance would be beneficial to the mining industry to address some issues relating to AMD and would help integrate environmental and economic considerations into sustainable environmental management.

Jan 1, 2002

By C Kusi-Manu, E Nyamusika

Geita Gold Mining Limited (GGML) is owned and jointly managed by Ashanti Goldfields Company (Ghana) and Anglogold (South Africa). The mine is currently one of the largest gold producing open pit operations in Africa producing 620 000 ounces of gold and mining in excess of 70 000 000 tonnes of material per annum. The mine currently has 2000 full time employeeÆs and a mine life in excess of 12 years. Geita Gold Mine is comprised of 11 different gold deposits situated in the Lake Victoria Goldfields of North Western Tanzania (Figure 1). The gold deposits are primarily associated with banded ironstone formations within an Archaean greenstone belt. Current global resources, including Inferred, stand at 129.5 million tonnes grading 3.91 g/t (16.2 M oz). Open pit reserves total 70.4 million tonnes grading 4.16 g/t (9.4 M oz). The majority of the gold resource is hosted by the three deposits (Nyankanga, Lone Cone and Geita Hill) within the æGeita TrendÆ a large and continuous (>4 km long) mineralised zone (Adofo, 2002). Geita Gold Mine commenced operations in June 2000, as the second large-scale open pit gold mine in Tanzania. The new workforce of 800 people had no previous experience in open pit gold mining. Since the start of the operation, Geita Gold MineÆs management ambition was to become a world leader in the application of best practice in all technical disciplines and in operational success. Staff development and training was identified as one of the major components in the company achieving its aim. Geita Gold MineÆs management has committed the necessary training staff, resources and the commitment of over $US 1 M a year to staff training and development programs for 2003 and 2004. The process and management of Geita Gold MineÆs mining geology staff training and development program is presented in this paper.

Jan 1, 2003

By K Li, C Zhang, i, J Oh, S Saydam, H L. Ram

Rock surface roughness plays an important role in excavation engineering projects and applications related to surface support structures, such as shotcrete and thin-sprayed liner (TSL). The morphology of the substrate affects the mechanical behaviour at the interface significantly. Therefore, it is necessary to quantify the roughness to investigate the relationship between roughness and mechanical properties. The joint roughness coefficient (JRC) proposed by Barton and Choubey (1977) is widely used in practice. It initially estimated the roughness by comparing it to 10 standard profiles visually. However, this method has lots of limitations, it is not only subjective but also a two-dimensional parameter with a small lab scale (Beer, 2002). The subsequent research on JRC, from visual comparison with Barton’s standard profiles to measuring the roughness with a profilometer (Weissbach, 1978; Alameda-Hernández et al, 2014). Then, 2D profiles or 3D surfaces by different technologies with higher resolution and accuracy, such as 3D laser or optical scanners (Fardin, 2004; Jiang, 2020, 2021). These 3D scanners are expensive and only suitable for laboratory scales. LiDAR scanners (Lato, 2009; Aubertin, 2022) have been applied in industrial size, but it is difficult to achieve a millimetre-level accuracy, which is mostly used to detect the thickness of shotcrete in practice. Thus, finding a cost-effective surface roughness measurement method with high accuracy is necessary. Photogrammetry is another method to digitalise the roughness of the surface. Similar to 3D scanners, photogrammetry also generates the point cloud of the objective surface and then reconstructs the 3D model of the objects. 3D reconstruction technology is a trendy technology in recent years. The structure from motion (SfM) algorithm (Westoby, 2012) to rebuild the 3D model with a flexible workflow, aligning photos, building dense point cloud, mesh and texture, which has been proved that can rebuild an accurate model in variable fields.

Nov 29, 2022

The Callie Underground Mine is located in the Tanami Desert 560 km Northwest of Alice Springs in the Northern Territory and is currently operated by Newmont Australia Limited with all mining activities carried out by Henry Walker Eltin. Callie was discovered in the region in 1991 with the box cut portal commenced south of the pit in 1995 and underground ore production commenced in 1998, after completion of the profitable Callie Open Pit. Development is currently occurring at depths of over 700 m with the decline currently below the 640 Reduced Level (RL) level (surface RL of 1400) with production haulage from 600 m below the surface. The combination of auto compression, a moderate geothermal gradient of 2¦C/100 m and diesel heat transfer causes the wet bulb temperature of intake air at the bottom of the mine to be typically 5¦C higher than on the surface. Wet Bulb temperatures can be up 5¦C higher again in active areas under secondary ventilation. During the 2002/2003 wet season many activities in the lower levels of the mine were stopped due to miners experiencing adverse heat related symptoms or the temperatures exceeding the Underground Heat Management Procedure stop job criteria of 32¦C wet bulb. To improve the working environment, combat the heat related illness and improve operational efficiency, a temporary 1.8 MW refrigerated cooling plant consisting of three 600 kW modules were installed underground to provide bulk air cooling for the lower levels of the mine during the 2003/2004 wet season. Several studies have been conducted in relation to future ventilation demands at Callie and all came to the conclusion, that some mine cooling will be necessary for Callie to meet the high production and development demands. This paper will describe the ventilation network at the Callie Underground operation focusing on the installation, operation and effectiveness of the temporary underground refrigeration plant.

Jan 1, 2005

By E R. Ramanaidou

The presence of clays in channel iron deposits (CID) within the Pilbara region of Western Australia causes problems in the classification and subsequent processing of the iron ore. In particular, the presence of clay adversely affects the processing and ore metallurgical properties by increasing the amount of Si and Al. Iron ore companies are seeking new methods to quickly and objectively detect the type and amount of clay in these deposits. Traditional methods of clay quantification such as X-ray diffraction (XRD) and thermal analysis are time consuming and expensive. Reliance on bulk chemical techniques (eg X-ray fluorescence, (XRF)) to derive normative mineral abundances can give misleading results. For example, the replacement of Fe by Al in the structure of the iron oxide minerals, goethite and haematite, which dominate in CIDs, can be significant. Silicon can also occur in association with these Fe minerals. If not accounted for, chemical techniques may overestimate the normative abundance of clays. In comparison, reflectance spectroscopy in the visible-to-near infrared (VNIR) and short wave infrared (SWIR) wavelength range (eg 400 - 2500 nm) and Fourier transform infrared (FTIR) (6000 - 900 cm-1) analysis are non-destructive techniques that can provide rapid and objective mineralogical information. Both techniques are particularly suited to Al-, Fe- and Mg-OH bearing minerals such as clays. Kaolinite and, to a lesser extent, montmorillonite are the most common clay contaminants that occur in the Pilbara CIDs. Previous applications of VNIR-SWIR spectroscopy have demonstrated the potential of these methods in quantifying the clay mineralogy of CID ore. In this study, variations in the depth of diagnostic spectral adsorption features in the SWIR wavelength range of kaolinite and montmorillonite in standardised mixtures, with known proportions of clay, were well correlated to the abundance of kaolinite and montmorillonite, and to the clay alumina content (ie per cent Al2O3 û clay).

Jan 1, 2007

ON milling plants, where a large number of slime samples are collected each day, considerable work is involved preparing them for the assayer. The usual practice (as the material usually passes a 150-mesh screen) is to place the sample on a 30-mesh screen an 1 brush it through on to a cloth below. The sample on the cloth is mixed and a portion taken out for the assayer. A simple apparatus that would rapidly disintegrate and prepare these slime samples for the assayer was needed. The manufacturers of assayers' supplies appear to have overlooked apparatus for this class of work. To overcome the tedious preparation of slime samples by the above method, the writer devised the apparatus shown in the accompanying drawing (Fig. 1). It will be seen that the apparatus is a combination of a modified household flour sifter and a riffle sampler.The apparatus is made up in two parts, the top half consisting of a screen box carrying the disintegrator, and the bottom half a, riffle divider, the joints between the two being a slip-in fit. A lid is provided for the screen box. The disintegrator consists of an open drum, to which are soldered three rows of spring steel fingers. Each row has a number of steel fingers which, on rotation, press against the semi-circular screen in the screen box, forcing the slime particles through.To operate the apparatus an 8-in. sample tin is placed under each spout of the divider. The sample of slime, which is about 2 lb. in weight and usually in lumps, is placed in the screen box, the cover replaced, and the drum rotated about half-a-dozen times by means of the handle.The disintegrated sample is divided by the rifle divider into two portions, which are found in the tins under the spouts of the apparatus. One of these samples is taken, mixed on a sheet, and sent to the assayer.

Jan 1, 1922

By S Ellis, J McMullen, M Habner

A new leaching regime was implemented at Kalgoorlie Consolidated Gold MinesÆ (KCGM) Gidji plant on a cyanide leach of ultra-fine ground sulfide concentrate. High lime concentration and lead nitrate were used to overcome the refractoriness of the fine sulfides (high reactivity of fine sulfides with the presence of gold tellurides). The leach pretreatment stage was shown to be detrimental and was shut down. Lime titration was introduced as part of the process control. It was shown that the leach could be performed at low dissolved oxygen, (ie 3 - 5 ppm) to avoid sulfide oxidation and the associated high cyanide and lime consumption. Plant results showed reduced cyanide consumption (40 per cent) and an accelerated leaching rate, whereby most of the gold was dissolved in 24 hours compared to 72 hours used previously. Similar work was performed at KCGMÆs Fimiston plant, which leached a combined ultra-fine ground sulfide concentrate along with the slimes fraction from hydrocycloning the bulk flotation concentrate. The use of lead nitrate, high lime and carbon-in-leach generated a faster leaching rate and minimised adsorption of gold on fine pyrite particles. The grind, previously controlled with a P80, was replaced by a P98 to give a better operational indication of top size hydrocyclone performance. Before the implementation of the changes, average leach residues for slimes and UFG were respectively 4.0 g/t and 5.5 g/t. With the modifications to the leaching regime, average leach residues were lower by 1.5 - 2.0 g/t. Whilst significant improvements have been demonstrated, further work is progressing at KCGM to further reduce ultra-fine concentrate leach tail grades, and also to improve the understanding of concentrate refractoriness and the factors responsible for differences between laboratory and plant leach results. Estimated profits since the implementation of the new approach are in the range of $A3.5 million per year.

Jan 1, 2005

Kumba Iron OreÆs Sishen Mine is one of the worldÆs largest iron ore mining operations situated in the Northern Cape, South Africa. It produces 29 million tonnes of iron ore products per annum using a dense medium separation (DMS) beneficiation process. Kumba is currently commissioning the Sishen Expansion Project (SEP) which will produce a further 13 Mtpa product through a jigging beneficiation process. The products from the Sishen Mine are highly sought after and consumed by domestic and various European and Asian customers. During beneficiation iron ore slimes are generated and discarded as waste material to the slimes dams. A study was initiated in 2005 to investigate processes to beneficiate these slimes to saleable products û prompted by the current high demand for iron ore products and the high grade in the Sishen slimes thickener underflow of approximately 54 per cent Fe. Initial laboratory scale test work indicated that a 66.5 per cent Fe iron ore concentrate could be produced from the Sishen slimes thickener underflow through a process of classification, two stage wet high intensity magnetic separation (WHIMS) and five stage flotation. This process, however, did not produce the desired results during pilot plant testing and numerous alternative process routes were investigated. The process that did achieve the desired product grade of 66.5 per cent Fe at economically feasible product yields consisted of two stage WHIMS using SLON magnetic separators. Results achieved indicate that the thickener underflow slimes can be upgraded from 54 per cent Fe to 66.5 per cent Fe at a product yield of ca. 40 per cent. Product yield could be increased through the further beneficiation of the SLON tails. The Sishen slimes thickener underflow production rate is ca. 3 Mtpa û resulting in a potential concentrate production of ca. 1.2 Mtpa. The slimes concentrate will be fed to a pelletising plant that will produce 66 per cent Fe pellets destined for domestic and or export markets.

Jan 1, 2008

By R Dimitrakopoulos, I Minniakhmetov

Geostatistical simulation techniques are used to quantify uncertainty of spatial attributes of interest describing mineral deposits, petroleum reservoirs, hydrogeological horizons, environmental contaminants and so on. The majority of existing methods consider second-order spatial statistics and Gaussian processes, while the more advanced multiple point-based simulation approaches are algorithmic and do not consistently account for the high-order spatial relations in data. Recently, simulation techniques for complex and non-Gaussian spatially distributed variables have been developed, based on high-order spatial cumulants, and make no assumptions on data distribution or require data transformations. In this paper, the previous developments are extended and a new approach for the joint simulation of multiple correlated variables using high-order spatial statistics is proposed. The technique is based on a new algorithm for the decorrelation of correlated variables into factors, using the so-termed diagonal domination condition of high-order cumulants. The decorrelated factors are then simulated using high-order simulation and back-transformed into the initial correlated variables. The decorrelation using diagonal domination of high-order statistics is tested with a data set from a multielement iron ore deposit, and then a fully known two-dimensional data set with two correlated variables is used to demonstrate the practical intricacies of the proposed method.CITATION:Minniakhmetov, I and Dimitrakopoulos, R, 2014. Joint simulation of correlated variables using high-order spatial statistics for orebody modelling, in Proceedings Orebody Modelling and Strategic Mine Planning Symposium 2014 , pp 53–60 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 24, 2014

As a result of cheap Maui gas and plentiful hydro-electricity, over the past few decades New Zealand has enjoyed some of the cheapest energy of any OECD country. The situation is changing; at best we are facing some radical changes in energy use, at worst we are facing a crisis resulting from the following:   Cause 1: Maui gas will be exhausted by 2012. Distorted pricing and supply contracts; resulted in Huntly Power Station, a coal mine mouth generator running on 70 per cent Maui gas until very recently!   Cause 2: Investment in new generating capacity (and transmission) is woefully inadequate; as a result hydro lake levels are being continually pushed to the limits.   Cause 3: Distorted thinking is everywhere, for example; æWe will find more cheap gasÆ. æGas is clean û coal is dirtyÆ. Hot air  û ægreen alternatives such as wind generation will save usÆ. æNo NuclearÆ. The outcomes of the distorted thinking will be price increases, then shortages (there is nothing like "black-outs" to focus national thinking) and finally recession.   But what can NZ do? - Mine Coal. Coal jump-started the industrial revolution, it can provide the transition to post-gas and oil. NZ has the second highest reserves of coal per capita in the world, yet we use less that any other country with significant coal reserves. Our exports of our best coal have increased threefold, yet out own consumption has remained static at about 1-1.5 Mtpa for the last 30 years. Our coal is NZ owned and inexpensive relative to other energy sources, even allowing for infrastructure and environmental costs.   What can we do? We as mining professionals can promote informed responsible development and investment in the country's mining and energy business. We can galvanise public opinion on issues such as æunbundlingÆ our entire energy chain from mine-to generator-through transmission, even if the nuclear issue has to wait! With oil above US$70/bbl substitution is feasible. For individuals, there is always solar!

Jan 1, 2006

By Mackenzie W. R, Harding C. G

The Islamic Republic of Mauritania is a large country of over 1 000 000 square kilometres situated on the Atlantic coast of north-west Africa. It is bounded to the north by Western Sahara and Algeria, to the east by Mali, and to the south by Mali and Senegal. With the exception of a narrow strip in the south along the Senegal River valley, the country lies entirely within the Sahara. The north of the country is largely devoid of vegetation, and it is estimated that 90 per cent of the population of 2.1 million people live in the southern quarter of the country. From about the turn of the century, Mauritania was ruled by the French from Dakar in Senegal and, until 1960 when it obtained independence from the French, Mauritania did not have a capital. Today, the capital is Nouakchott, home to an estimated 600 000 Moors of mixed Arab, Berber and African descent. After independence in 1960, disputes over Spanish Sahara to the north led to a military coup in 1978 and it was not until 1991 that a democratic constitution was adopted. The new constitution declares Mauritania to be an `Islamic, African, and Arab Republic.' It provides for an executive president, elected for a six-year term, and for a bicameral legislature, consisting of a national assembly and a senate. The first presidential and legislative elections were held in 1992 under United Nations' supervision. Mauritania's economy is predominantly pastoral, and the country depends heavily on foreign aid. Its principal export industries are iron ore mining (about ten million tonnes per annum from mines near Fderik in the north) and fishing. Mines d'Or d'Akjoujt SA (MORAK), a company in which General Gold Resources NL (GGR) has 42.5 per cent equity, produces over 50 000 ounces of gold per annum, which represents the third largest export industry in the country.

Jan 1, 1996

Segregation of granular material is a very common phenomenon. Realised in a forced manner it is a basis of all separation processes in mineral processing, mainly classification and beneficiation. Spontaneous segregation, occurring under dynamic conditions and making possible mutual translocations of grains, is however an undesired phenomenon as a rule. It decreases the degree of blending of the material which in many cases is an important technological feature of the material. It should be noted that the segregation can be treated as the reverse phenomenon to blending. The importance and frequency of spontaneous segregation has caused many rules and technical solutions to be developed aimed at decreasing the possibility of the spontaneous segregation. However, not enough attention has been focused on an quantitative description of the phenomenon of segregation, either to ways of quantitatively evaluating segregation effects or to explanation of the influence of conditions under which the grained material exists on the occurrence of segregation and its degree. The paper presents the following topics: 1. new numerical indicator for evaluation of the segregation degree (and blending) of material contained in an arbitrary determined region is proposed; it is determined relative to the maximum possible degree of segregation of the given material when taking into account its individual properties; values of the indicator can be calculated effectively for an arbitrary material, also for ulticomponent material. 2. there is proposed a way of evaluating the state of segregation (or blending) of material in an arbitrary part of the region mentioned above, based on investigation of the gradient of distribution of the material features; algorithms of analytical procedures and calculations are presented for typical cases met in mineral processing; 3. results of initial experimental studies, both laboratory and industrial, concerning the effect of the chosen factors occurring in transport and storage of the material on its degree of segregation are presented; and 4. conclusions referring to the fundamental and practical significance of the results are also presented.

Jan 1, 1993