Search Documents

By B Mil, D P. Adhikary, M S. Craig, P Rutzou

As shallow coal reserves with easier mining conditions are exhausted, future deep coal mining with difficult mining conditions may face significant economic, technical and environmental challenges. Moreover, due to increased coal production capacity with the new generation longwall systems, predicting adverse mining conditions has become more important than ever before. The potential impact of these difficult conditions is an increase in mining risks and costs. Reliable prediction of rock mass deformation, mine stability and mine water inflow is not only essential for improving mine safety and reduction of coal production costs, but it is also important for the assessment and management of the environmental impact of mining. This paper briefly describes an integrated approach to mine groundwater simulation and prediction. Central to this integrated approach is a three-dimensional computer code called COSFLOW, developed by CSIRO of Australia in collaboration with NEDO and JCOAL of Japan to address the pressing issues of mine water management. COSFLOW incorporates unique features that make it ideal for the simulation of coal mining and examining the interaction between rock fracture, aquifer interference and water flow. This paper also presents an example of an application of the integrated approach and COSFLOW in water inflow prediction at Springvale Mine in New South Wales.

Jan 1, 2006

By Munive L

As a result of commissioning No.5 B F, 2670m3 inner volume, in March 1972, the largest B F in Australia; new expertise in maintenance, furnace operating practices and design have been obtained. This experience has also been extended to the other furnaces. A series of operating problems were con- fronted; all parts of the furnace, including the hearth side walls and bottom,were adversely affected. In December 1977, serious hearth wall wear was detected at No.5 B F, and a potential hearth breakout was averted. Hearth wear in No.4 B F was discovered in March 1979, and emer- gency underhearth cooling was installed to reverse an extremely advanced hearth bottom erosion. Hearth behaviour of No.5 B F oampaign 1972-1978 and`No.4 B F campaign commencing 1972, has been studied and discussions on the causes, effects and measures undertaken to protect the hearths are presented in this paper.

Jan 1, 1981

Low-grade manganese ore from Sandur, India containing 28.2 per cent manganese and 20.9 per cent iron was roasted with sulphating agents, viz ammonium sulphate and sulphuric acid. The effects of temperature of roasting (300 - 800°C), time (15 - 120 minutes) and amount of sulphating agent (0.5 to 2.5 times stoichiometric amount) were studied. Under optimum conditions, it was possible to extract 92 per cent manganese with 0.5 per cent iron at a temperature of 700°C for 90 minutes duration with stoichiometric ammonium sulphate. Further increase in quantity of ammonium sulphate has marginal effect on manganese extraction.Similar studies were carried out with sulphuric acid, and an extraction of about 87 per cent manganese with similar iron contamination was achieved at a temperature of 700°C for 60 minutes duration with stoichiometric quantity of the reagent. Selective extraction of manganese with respect to iron is discussed.

Jan 1, 1996

By Toop G. W, Martin M. T

At the Cominco Ltd plant in Trail, British Columbia, the Betts lead electrorefining operation requires that antimony and arsenic be present in the lead bullion for anode slimes stability control. Until 1986 the antimony and arsenic inputs from concentrates to the Lead Smelter needed to be supplemented with high antimony and arsenic content recycles from the Silver Refinery operation to maintain a satisfactory composition. Subsequently the antimony and arsenic inputs have increased as a result of increasing levels of antimony in Sullivan lead concentrate and the purchase of larger quantities of silver/gold bearing concentrates that inherently have higher antimony and arsenic levels. The increase has exceeded the limits required by the Betts process. It therefore became necessary to install a partial softening process downstream of the Continuous Drossing Furnace to satisfactorily control the bullion composition. Process control and oxygen injection techniques used in the partial softening process are reviewed. Approximately 8 tonnes/day of softening slag are produced having a combined antimony and arsenic content of 25 percent. This slag is reduced in a recently commissioned short rotary furnace followed by vacuum distillation of the metal for recovery of the antimony and arsenic values.

Jan 1, 1989

This paper describes a case history of using an integrated geochemical and hydrogeological approach to predicting solute transport from gold mine tailings in highly fractured volcanic terrain. Two modes of transport are identified, namely discrete fissure flow and diffuse seepage through the intervening weathered rock. A simple mass balance approach is used to predict concentrations at points of emergence because most of the seepage migration will be by way of the discrete fissures where advective transport will predominate. The mass balance method has been modified to allow for natural degradation of weak acid dissociable cyanide. The long term effects of diffuse seepage are also examined.

Jan 1, 1988

By Corrans I. J

A wide variety of copper-bearing gold ores exist throughout the world which have yet to be processed economically for gold recovery. The copper minerals present in these ores can be many and varied and usually result in high cyanide consumptions and problematical levels of copper in leach solutions. The copper content is often assessed as being of nuisance value as the ores are generally of the grade or type which can support neither a copper project nor a straight gold project, due to the interference of copper. Many of these problematical gold-copper ores could possibly be processed economically for gold recovery using a novel ammoniacal cyanide leaching system which utilises a copper-ammonia-cyanide complex as an effective lixiviant for gold, yet limits copper dissolution and conserves cyanide. Results from recent testing have promoted a re-evaluation of the economics of treating some gold-copper ores. This paper describes the novel leaching process and discusses the conditions employed to treat a variety of gold-copper ores.

Jan 1, 1992

By McDonough RJ

Mechanical agitation is the preferred system for Carbon-in-Pulp processes. Agitator design requires an understanding of the relationships between energy, shear and flow and the nature of shear rates and shear stress distribution in an agitated tank. Laser techniques are useful when studying these phenomena. The alternatives are open impeller agitation or draft tube circulation. Each has advantages and disadvantages but both require careful choice of impeller and system geometry. Overall the draft tube circulator is the more attractive in many cases. Open impeller agitators are designed in accordance with established procedures for mineral slurries taking due account of shear effects in tanks containing carbon. Preferred tank proportions are 1:1. Circulator design methods were developed in the alumina industry before being applied to CIP processes. Optimum draft tube diameter is critical to power requirements. The overall system must be designed as an axial flow pump operating in a closed circuit. Such a system exhibits quite characteristic behaviour which requires an impeller of varying pitch angle and aerofoil section for best stability and efficiency.' Tank height may be 2 or more times diameter. Resuspension of settled solids can be assisted by a series of slots in the draft tube. These slots have the secondary effect of stabilizing hydraulic behaviour and reducing installed power requirements.

Jan 1, 1982

The basic methods to remove material from underground mines are to: convey continuously, batch in trucks, or batch in a shaft hoisting system. A critical aspect in shaft hoisting is accurately batch loading the conveyance or skip. During the 20th century, the almost universally accepted method of loading was via a measuring flask. A measured quantity of material was filled into the flask as the conveyance traveled through the shaft and then, when the conveyance arrived at the loading position, it was ædumpedÆ into the skip for hoisting. Two skip loading conveyor systems have been developed in Australia: BHP BillitonÆs Cannington Hoisting System, installed in 1996, is believed to be the first conveyor loading system designed for two skips. Two parallel conveyors were installed with a two-way gate controlling feed onto each belt. Cannington was ideally suited to conveyor loading as the very high ore density and relatively low production rate required a relatively small volumetric storage capacity on the weighed belt. Newcrest MiningÆs Telfer Hoisting System, which is to be commissioned in 2005, consists of a single slewing conveyor loading each skip with 32 tonnes of ore. The development of this system eliminates the two-way transfer, improves the accessibility and minimises the excavation required. While every hoisting system must be considered on its merits, conveyor loading offers significant advantages to the Mining Industry in the 21st century.

Jan 1, 2005

By Newell R

The situation in which a gas stream flows through a bed of solids is found in many extractive metallurgy operations. Examples include blast furnace smelting, sintering and pellet firing. In such cases the resistance of the bed to fluid flow and heat transfer is of major significance to the operation of the process. This depends on the structure of the bed, which in turn is influenced by the manner in which the materials are charged to the bed. In this work, packed beds of hematite pellets of two different sizes were prepared in which the pellets were charged as a uniform mixture and in layers. The porosity, flow resistance and heat front velocity were measured and compared for the different bed configurations. For layered beds, the overall porosity and heat front velocity increased while the flow resistance decreased when compared with the values obtained for the uniform mixture. The size of the layers also had an effect on flow Properties.

Jan 1, 1991

By Takeuchi T, Matsugaki K

We have developed a new shotcreting which is suitable for the primary support of TBM tun- nelling in soft rock. This spraying compound is made from cement, a special accelerating agent and glass fiber. It can be sprayed on to a rock surface using a specially designed gun similar to paint-spraying equipment. Some of the characteristics of this method are as follows: 1) If spraying is done keeping the distance between the gun and the rock surface be- tween 0.2 and 3.0 m, the rebound losses are small and there is little dust. 2) This method can be used on excavated rock surfaces under inflowing water conditions of up to 10 i/m2òmin. 3) After 1 day the compressive strengthes of the sprayed material is greater than 250 kgf/cm2.

Jan 1, 1987

By Dredge K, Becker G

The successful dissolution of the valuable metal components of a solid phase is often a basic requirement for a successful hydrometallurgical treatment process. Sulphide concentrates are often an intermediate product in the production of refined metals. These concentrates need to be further processed to realise the contained metal values in a useful form. Pyrometallurgical techniques have tended to dominate, at least the initial treatment of metal sulphide concentrates. The technology described, Activox«, has been developed specifically for the hydrometallurgical treatment of these sulphide mineral concentrates, in particular those containing copper, nickel or gold. The Activox« dissolution technology coupled with the use of existing, or newly developed, downstream processing techniques, offers alternative processing options for the recovery of full value, saleable metal products. The technology is suitable for application down to relatively small scales opening up the opportunity for on-site production of metals even at modest scale mining Operations.

Jan 1, 1995

By Morrison RD, Peberdy NJ

Simulator training provides a cost effective method of training operators in both the process and control aspects of a metallurgical plant. The components and requirements of a training simulator are described along with the benefits to be gained from simulator training. A table summarizing some applications of metallurgical training simulators is Given.

Jan 1, 1991

By Tumilty JA, Douglas WD

With a view to characterising activated carbons with respect to their ability to adsorb gold from cyanide solutions, an attempt was made to establish correlations between the equilibrium capacity constant of the activated carbon for gold (k value) and five other parameters. These parameters were percentage ash content, iodine number, methylene blue value, surface area, and pH respectively. Twenty four different proprietary brand activated carbon products were examined, the correlations being established by means of linear regression analysis. It was shown that a significant correlation only existed between k and pH value. This was interpreted in terms of the presence of oxide complexes on the activated carbon surface. Further test specifications as developed at the Anglo American Research Laboratories are also outlined with a view to the selection of activated carbons suitable for use in' a carbon-in-pulp operation. Methods for assessing both the capacity and the kinetic response of carbons for gold adsorption are presented, while the measurement of attrition resistance, particle size distribution and moisture and solubles content is also outlined. The measurement of the capacity constant for gold adsorption together with attrition resistance are considered to be necessary in the initial selection of a suitable carbon.

Jan 1, 1982

By Drew GJ

Smelting is a process in which heat is applied to ores to produce metals. Impurities are removed by segregation in the slag, or converted to gases or sublimates and discharged up the flues. Mate- rials delivered into the smelter have generally been previously concentrated or processed at the mine, and the size of their particles, and also their chem- ical composition, greatly affect the smelting proc- ess. Some impurities are more difficult to remove than others, and are more objectionable in the finished product. The first metal mine in Australia was opened near Glen Osmond in South Australia in 1841, for silver-lead ore. It was followed by the Kapunda Mine for copper in 1844 and by the very signifi- cant Burra Mine, also for copper, in 1845 (Figure 1). A major part of their ores could not be exported profitably for smelting in England and Wales where the large smelters were then located. In addition to treating ores from Cornwall these also smelted ores from recently opened mines in Cuba, Chile, and Mexico. The Welsh smelting process (Figure 2) was developed over a period of years to utilise Welsh coal to treat copper ores particularly from Corn- wall. The first step consisted of a calcination to remove sufficient sulphur from sulphide ores so that subsequent melting operations would produce copper matte, a copper-iron-sulphide which melted at about 1050¦C. Subsequently refined to white metal with the addition of carbonate and oxide ores when these were available, and often through an intermediate step of blue metal, this matte was finally oxidised to pure copper, or as pure as it could be made in those days. The movement of metal towards greater purity was matched by a reverse flow of slag, carrying reducing quantities of copper, so that the slag rejected from the first melting process carried the minimum quantity of copper. Chemical analysis was virtually unknown. Materials were assayed in the laboratory by a small scale smelting process, and the purity of the final refined copper was judged b} the ease with Mhich it could be hammered into thin sheets.

Jan 1, 1987

We operate a 50 cubic metre per hour mobile alluvial tin mine on small scattered deposits in the Irvinebank area. Wash is loaded by a Cat. 950 front end loader into the bin of a screening section and screen undersize is pumped to a separate jig unit. Both units are rubber tyre mounted and are shifted once a month when the haul distance gets to about 150 metres. The shift takes about 8 hours.The screening section is sited close to the face with a ramp for dumping in and a one metre drop for the screen oversize which is removed by the loader.A reciprocating plate feeder with variable rate hydraulic drive, feeds wash onto a horizontal 1.2 metre x 2.2 metre two deck Kinergy screen. Minus 10 millimetre undersize is pumped by a Warman 6/4 sand pump with a 37 kW variable speed Reeves Motordrive to the jig unit.The jigs are placed 20 to 70 metres away so that tailings will run back into previously mined ground. Reject stone is used to impound the sand and mud, allowing the water to filter through and return to the feed water pump.An MTE trapezoidal module is used as a primary jig. Hutch product is pumped to an eight cell clean-up jig which we developed ourselves along MTE lines. It has a trapezoidal shape and sawtooth stroke provided by cams. Clean tin is taken from the first four spigots. These consist of clear plastic tubes fitted with taps so that there is no water loss except when the tin is drawn off. Fine concentrate from the last four spigots is put over a curvilinear table which sits on a steel base packed with earth.

Jan 1, 1986

By Braaksma M, Dudley K

The Girilambone Copper Mine is located in central New South Wales and has been operational since May 1993. Girilambone is operated by the Girilambone Copper Company (GCC) which is a joint venture between Straits Resources (60 per cent) and Nord Resources (40 per cent). The mine complex consists of an open cut pit, ore crushing and stacking, heap leaching, solvent extraction (SX) and electrowinning (EW) allowing for the recovery of 14 600 tpa of LME grade A cathode copper. The leaching and SX-EW plant have been described previously by Rear, Taylor and Busby (1994) and Miller (1994).

Jan 1, 1996

Kambalda Nickel Mines (KNM) currently have eight operating underground mines, three on care and maintenance and other orebodies at various stages of development. Each operation differs greatly in scale, ore grade, milling characteristics and cost structure. The optimum number of mines operating at any one time and their level of production, within the constraints of mill blending requirements and overall mill capacity, is the key to maximising profit in a climate of fluctuating nickel price. Cost models are developed for each mining operation which relate ore tonnes mined to total cost per pound of nickel produced from that mine. Aggregating the mine cost models provides a composite cost model for the entire operation. The cost models provide the operating levels necessary for each of the mines and hence Kambalda as a whole, to operate at the lowest cost per pound of nickel produced. However, lowest operating cost does not necessarily lead to maximum profit. The level of nickel price above operating cost can dictate maximum profit positions which are contrary to the lowest cost position. In addition, the decision to temporarily shutdown mines or reopen mines in response to nickel price movements comes at a cost. In considering the optimum operating position, the price of production capacity flexibility needs to be taken into account. This paper describes the process used to assess all these factors in arriving at the optimum operating position for Kambalda Nickel Mines for varying nickel prices. The same methodology can be used by corporate management in managing multiple mine sites as well as mine management managing multiple orebodies or stoping blocks for maximum profit.

Jan 1, 1995

Trial ore production began at the Hilton Mine in 1987 with ore being trucked into Mount Isa for treatment at the Lead/zinc Concentrator. The expansion of output from the Hilton mine required additional concentrator capacity which led to the design and construction of the Hilton Concentrator. Systematic laboratory and Pilot Plant testwork dating back to 1968 to develop a flowsheet and reagent regime suited to the varied and generally complex mineralogy of the Hilton ore was successfully concluded with a Pilot Plant trial using a sulphoxy based sphalerite depression system in early 1988.In May 1988, a decision was made to build a 750 000 t/y concentrator with some novel technological and capital-efficient features. The Hilton Concentrator commenced operations in December 1989 reaching target grinding rates and metallurgical performance by mid-1990. Current ore treatment rate is 800 000 t/y.Lead concentrates assaying 43 - 48 per cent Pb are combined with the Lead/Zinc Concentrator lead concentrates and smelted at Mount Isa. Zinc and Low Grade Middlings (LGM) concentrates are filtered in the zinc filter plant at Isa with those from the Lead/Zinc Concentrator and then railed to Townsville for export or treatment in Australia.

Jan 1, 1993

The Mt Leyshon Volcanic Complex lies 24 kilometres south-west of Charters Towers and is part of a major continental volcano-plutonic province of Permo-Carboniferous age in north-east Queensland.The volcanic complex is located on the contact between Cape River Beds of Cambrian age and the Ordovician Ravenswood Granodiorite Complex. The emplacement of the complex appears to have been structurally controlled and is at the intersection of major north-east and minor north-west trending lineaments. The gold mineralisation occurs in the form of cavity coatings, matrix replacement, disseminations, and vein and fracture fillings containing quartz, chlorite, carbonate, pyrite, sphalerite, chalcopyrite and galena in various volcanic breccias and rhyolitic intrusives. The current published mineable ore reserves, calculated by Noranda Australia Limited, are 6.3 million tonnes of 1.8 grams per tonne gold and 4.8 grams per tonne silver. The orebody is open at depth, and in all directions except north where it abuts the granodiorite.

Jan 1, 1986

By Lyman G. J

The volumetric distribution of valuable mineral, or phase, within a set of particles is called the liberation distribution. To optimise a mineral processing plant it is necessary to know the liberation distribution in most of the streams, especially those adjacent to separation devices. Possibly the most common method of analysing ore particles is to section sets of similarly-sized particles. However, the observed sections can appear liberated even if the particle is composite. This means that the 'apparent' liberation distribution can be very different to the 'actual' liberation distribution. This discrepancy is called stereological error. One must then somehow estimate the actual liberation distribution using the apparent liberation distribution. This problem is called the liberation distribution problem and its solution is the subject of this paper. The solution is based on a sorting algorithm which uses geometric probability equations. These equations are based on the assumption that sectioning is isotropic and uniform. Apart from these assumptions the equations are completely independent of particle shape or texture. The algorithm is verified by numerical simulation and experimental methods. For the numerical simulations the correspondence between the actual and estimated liberation distributions is excellent. For the experimental methods, the agreement is not as good and appears to be due to experimental errors.

Jan 1, 1995