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By Bratt G. C, Horsham T. R, Langlois D. R

Mercury is precipitated from weakly acidic sulphuric acid plant waste liquor at the Risdon Works of the Electrolytic Zinc Company of Australasia Limited. Factors determining the choice of process and reagent, as well as details of process development including scale- up, and the time scale of implementation are outlined. The plant is designed for a maximum flow of 120 m3/h.of liquor containing mercury and saturated with sulphur dioxide. Mercury is precipitated by fluidised beds of aluminium pellets contained in lead columns. Liquor is then clarified with plate and frame presses. Experimental work on the mercury precip- itation reaction was conducted in three phases involving a scale-up of 100 times in equipment size. Extensive pilot filtration work was carried out on a 1 m2 press. Scale-up of 120 times was required for specification of plant filters. The basis of scale-up is outlined.

Jan 1, 1977

By McGuckin P. H

The Cadjebut orebody is a MississippiValleyType (MVT) zinc-lead deposit, and is located approximately 80 km southeast of Fitzroy Crossing in the Kimberley region of Western Australia. It is one of five zinc rich orebodies discovered by BHP Minerals in the region.The participation in the Cadjebut Joint Venture is BHP Minerals limited (Manager), 58 per cent and Billiton Australia Limited, 42 per cent.The mineralisation is contained in gently dipping dolomitised limestone sediments (part of the Lennard Shelf carbonate sequence) with an orebody width of between 60 and 120 m and a strike length of approximately 3 km. The deposit consists of two thin, stratabound shallow dipping orebodies, one overlying the other, to provide discrete but not independent mining areas. The orebody contains approximately 2.1 million tonnes of recoverable ore at an approximate grade of 12 per cent zinc and four per cent Lead.

Jan 1, 1992

By Lawson F

Narrowly sized samples of synthetic chalcocite (Cu2S) and covellite (CuS) were leached in oxygenated acidic sulphate-chloride solutions. Chalcocite leaches via a two-stage process, the first stage being much more rapid than the second. The apparent activation energies are 42.7 kJ/mol and 68.2 kJ/mol respectively, indicating that both stages are controlled by surface chemical reactions. The leaching of covellite is also chemically rate limited with an apparent activation energy of 77.0 kJ/mol. The shrinking core model is able to be used to describe the leaching processes of both copper sulphides. Over 97% of copper present in chalcocite dissolved in 0.5 M H2SO4i0.5 M NaCl solution under one atmosphere pressure of oxygen at 85¦C in three hours from particles 36 pm in diameter while only 85% of the contained copper in covellite particles 13 pm in diameter dissolved in the same solution under the same pressure at 90¦C in six hours. About 99% of the sulphide sulphur in chalcocite was oxidised to elemental sulphur while only 90% of sulphide sulphur was converted to elemental sulphur during covellite leaching. The mechanism proposed for the two-stage leaching of chalcocite particles is as follows: in the first stage, cuprous ions rapidly diffuse to the external reaction surface where they lose electrons and go into solution as a complex cupric chloride ion CuCl'. The second leaching stage begins when covellite is formed at the external surface. Copper again goes into solution as the CuCl* ion. The mechanism for leaching synthetic covellite is similar to the second leaching stage for chalcocite except that the leaching rate of synthetic covellite is substantially lower than that of the newly formed 'covellite'. Elemental sulphur formed on the particle surface retards the dissolution processes during both chalcocite and covellite leaching. The presence of chloride ions is necessary if essentially all of the copper from chalcocite is to go into solution. When no chloride ions were present, the leaching rate of both newly formed and synthetic covellite was almost zero. The main role of chloride ions in both chalcocite and covellite leaching appears to be to disrupt the passivation film of sulphur on the particle surface.

Jan 1, 1989

By N Nosaka

Sinter plants currently have to use many types of ores with different properties due to the great demand for iron ores, which was driven by the massive growth in Chinese iron and steel industry. On the other hand, the development of iron ore mines is limited and the quality of iron ores is in a tendency of deterioration. Therefore, the quick evaluation of iron ores has become more important for a stable sintering operation. In addition to such a background, the CO2 gas emission should be decreased from sinter plants according to the Kyoto Protocol on reduction in greenhouse gases. The present paper first describes the methods for testing the metallurgical properties of iron ores, which use a small volume of sample mainly focusing on magnetite ore that is effective for reducing CO2 and NOx emissions from sinter plants. The sintering properties of ores are evaluated separately according to each stage in sintering process, that is: heating stage before melt formation, primary melt formation in adhering fines layer of pseudo-particles, and assimilation of nucleus ore particle with calcium-ferrite system melts.Next, the application of the tests to the blending ores is introduced showing examples of the studies on the effective usage of the magnetite ores with Marra Mamba type ores.

Jan 1, 2005

The excavation of sandstone of generally medium strength for the Liverpool Link Line by Dosco Mark 2A Roadheaders has demonstrated the versatility and good performance of road- headers in suitable rock formations. A description is given of the strata traversed by the tunnels together with descriptions of equipment and crews, excavation sequences, operating difficulties and the results achiev- ed. In addition tunnel construction is des- cribed for a difficult section where there was only O.Sm clearance between invert and a four lane vehicular tunnel below, and O.5m clearance between tunnel crown and a century old sewer above.

Jan 1, 1976

By Tamura S, Maekawa M, Nagai C, Deguchi M

Heavy fuel oil has been injected into blast furnaces in order to reduce the coke rate, to improve productivity and to control the furnace heat condition. However, in the first oil crisis at the end of 1973, it became necessary to develop an injection technique of new fuels into blast furnaces to reduce oil consumption. In Kobe Steel, Ltd., it was thought that the injection of a coal-oil mixture (COM) into blast furnaces could be developed in a shorter time than pulverised coal injection. The development of this technique was under- taken in 1974, and COM was injected through two tuyeres of the No.3 blast furnace at the Kobe works in 1977 and 1978. In addition, in order to study the combust- ibility of COM and the phenomena in the tuyere zone, the development of a gas sampling method was under- taken in 1975. In this paper, the results of the injection tests and gas sampling tests are reported.

Jan 1, 1981

At the Highway gold deposit, 37 kilometres south of Charters Towers, gold mineralisation occurs in several stratiform layers and is enriched near the surface by lateritic weathering. The host rocks of the mineralisation are ferruginous and baritic, intensely sericitised, kaolinised and silicified rhyolitic tuffs and pyroclastics of the Cambrian-Ordovician Mount Windsor Volcanics. The geology of the deposit is illustrated in Figures 1 and 2.Forty-six reverse circulation percussion drill holes were completed by North-Queensland Resources N.L. for a total of 2,093 metres of drilling to further outline and test the open pit resource. The drilling indicated a resource of 171,000 tonnes averaging 4.4 grams per tonne gold and 5.0 grams per tonne silver.A shallow initial pit, approximately 30 metres deep, has been designed to mine the high grade upper portion of the western stratiform layer and part of the eastern layer. This initial pit is designed to mine 45,400 tonnes of western mineralisation at an average grade of 7.84 grams per tonne gold and 22,200 tonnes of eastern mineralisation at an average grade of 1.42 grams per tonne gold. Waste is estimated at 110,900 tonnes.

Jan 1, 1986

By O C. Aitken

Mining project developers are committed to taking on all the challenges and obligations of financial risk management. It is interesting that our current models of project finance have not changed markedly in principle for some 700 years, yet the rules which prepare the way for effective project assessment and subsequent loan management still do not seem to be well understood by project or corporate developers. Probability and gravity fill the opposing trays of the golden balance of financial risk. Understanding the relationship between these opposing, but intimately related elements is critically important for successful risk management. It is important to move beyond the qualitative approach and use quantitative methodologies to measure the financial impact of the occurrence of each major risk. We live in a world where the only constant is change and thus it is not surprising that a bankerÆs life is still governed by assessing, measuring and pricing risk. Banks are attracted to funding well managed companies with strong cash flow projects. But we are more risk averse compared with the credit appetites of late last century. This attitudinal shift has resulted in tougher term sheets for the early 21st century, encompassing more conditions, warranties, undertakings and stronger cash sweep provisions. Banks require well balanced portfolios to achieve an acceptable spread of risk, thus there is a more restricted market for small company lending than that available to large multi-project corporate entities with less concentrated risk profiles. Because of this focus, junior company debt finance remains comparatively expensive and especially so for more challenging geographical and political environments.

Jan 1, 2004

During the last 12 years the dominant theme at Mount Isa has been production expansion. Production rate of 3,200 tons per mine day in 1953 increased to 11,000 tons per mine day in 1960, completing the first stage of the programme. The second stage is still in progress. It is planned to increase production of lead and copper ores to 19,500 tons per mine day in 1968.Current mine workings comprise an area 2 miles long and 1/2 mile wide, the deepest production level being 2,000 ft below surface. During the next 10 years it is anticipated that mining will be carried out to a depthof 3,200 ft below the surface. The rapid extension of the mine has necessitated frequent re-appraisals and modifications to the primary ventilation system, and a system with considerable in-built flexibility has gradually evolved. Since 1953, volume circulation has increased eight-fold to approximately 2,400,000 c.f.m.To suit various mine operating requirements, a combined pressure-exhaust system of ventilation was adopted.The usefulness of this system is at a maximum when the fans installed permit ready changes in the location of the boundary between adjacent spheres of fan influence.Air flow through openings to the surface can be controlled or used as required. This feature is considered necessary for a mine where underground workings are comparatively shallow and caving to the surface willoccur.A major heat problem exists due to climate conditions, high geothermic gradient (one degree Fahrenheit in 95 ft), and oxidation in leached and brecciated areas.Because of this, quantities of air far in excess of the minimum requirements for dust removal are required to ensure satisfactory environmental conditions. In assessing fan pressure, consideration was given to 1. the mine resistance, and the requirement for pressure differential capability,2. the pressure requirement for deeper mining, with due allowance for booster fan use where appropriate,and3. the desire to avoid fan operation at pressure duties above the region of peak fan efficiency.The latter feature provides a safety margin between maximum operating pressure and the stall point of the fan.

Jan 1, 1967

The BHP Billiton Cannington Mine uses paste as the principal method of backfill. The paste consists of mine tailings and cement. The pastefill process at Cannington includes a surface plant where the thickened total tailings stream is filtered using two vacuum disc filters operating in parallel. The filter cake is mixed with Portland cement in a spiral mixer to produce a paste at a specified yield stress and cement content. The initial system design only required paste to be delivered under gravity. In July 2001, Cannington Mine expanded to 1.8 million tonnes per annum. In order to achieve this, production commenced on five new levels underground, 350 metres horizontally from the paste plant. Due to the high viscosity of paste, this distance was too far to transport by gravity. Traditionally, paste has been transported only short distances (50 m) using positive displacement pumps, with very limited success. At Cannington, a conveyor was installed which was designed to achieve zero spillage into the environment. This paper discusses the innovative design strategies used to attain zero harm on the environment whilst conveying a material, which is difficult to transport.

Jan 1, 2002

By Fukuichi K, Matsukawa H, Nakagawa M

The Myojin Tunnel, located on the Shikoku Expressway being constructed by the Japan Highway Public Corporation, is 3,770 m long with a cross sectional area of 80 m2. Kajima Corporation, the contractor for the northern construction section of this tunnel, adopted here its newly developed "mini-bench" method of excavation. The objectives of this method were 1. to be able to perform all tunnelling operations using the whole cross sectional area of the tunnel under any ground conditions, and 2. to be able to vary the length of the mini-bench within the range of 0 to 7 m according to ground conditions. In order to achieve these objectives, original, sophisticated equipment was developed and employed in the work. One of the newly developed machines is a special jumbo, a hydraulic crawler jumbo with five rotary type booms. This machine is capable of accommodating variation in the length of the mini-bench accompanying changes in ground conditions and is equipped with a special mechanism by which the positions of the booms can be controlled by means of hydraulic cylinders. Also developed was a shotcrete robot for use with the mini-bench method which has the control board near the shotcrete nozzle for close control of application.

Jan 1, 1987

Despite the research interest on non-cyanide gold lixiviants over the past three decades all the non-cyanide gold processes are still at the developmental stages with limited understanding on solution and pulp chemistry in some cases. This is partly associated with:the difficulties in measuring reliable equilibrium data for various gold(I/III) complexes with non-cyanide ligands; the lack of knowledge on mixed-ligand complexes; and different kinetic stabilities of gold(I) complexes with respect to disproportionation. This paper presents reliable equilibrium data for complex formation, dissolution, precipitation, hydrolysis and disproportionation reactions of gold(I/III) compounds with a range of non-cyanide ligands: S2-, HS-, S2O32-, SO32-, SCN-, SC(NH2)2, OH-, NH3, Cl-. Solubility of gold(III) compounds with non-cyanide ligands follows the order: Cl- > NH3 > OH-. In the case of gold metal the solubility order at a given temperature is MgS > Fe(III) or Cu(II) / Cl- > Cu(II)/NH3 > NaOH, but the solubility increases with increasing temperature. Rate of gold oxidation with different non-cyanide lixiviants depends on the oxidant, ligand and temperature and follows the general order of oxidants: Fe(III) or Cu(II) > X2 or OX- > O2 (X = halogen). In the case of metal ion oxidants oxygen is required to reoxidise Cu(I) or Fe(II). Some of the reported data on chloride and thiosulfate leaching of gold are presented to highlight the effect of additives and the copper, silver and sulfur contents of the starting material on the success of non-cyanide leaching processes.

Jan 1, 2004

Since the announcement of early experimental work on rotating magnetic field (RMF) separation at the Sixth Mill OperatorsÆ Conference in 1997, the concept of RMF separation has been developed and tested further, and a range of separator designs is now being developed for the mineral processing industry. Shortened, but full diameter, prototypes of these designs are now available for test separations. The main advantage of RMF low and medium intensity attraction separation is the significant reduction in particle entrapment. This reduction is so significant that RMF separators are capable of operating very close to their theoretical limit for separator precision. This makes small laboratory RMF attraction separators ideal for low to medium intensity magnetic sample analysis, and larger separators ideal for applications requiring sharp magnetic susceptibility splits. Feedrates and separation rates are comparable to those of conventional attraction separators. Although RMF attraction separators are now well advanced in development, RMF rotation separation is still in the early experimental stage. The exception to this is the dry particle rotation eddy current separator that offers eddy current separation of small non-ferrous metallic particles from about 2 mm down to about 100 microns. Wet small particle eddy current separation is also possible, but has not yet been developed to a practical stage.

Jan 1, 2000

The Chamber of Minerals and Energy of Western Australia Inc (CME) is the peak resources sector representative body in Western Australia. The paper gives a summary of the role and objectives of the CME, including its history, structure and key program areas. Particular attention is given to the role of the CME in the Eastern Goldfields of Western Australia. The paper details the types of CME membership that are currently in existence. The paper argues that a strong and vibrant CME is vital to secure the resources sectorÆs growth and development in future years.

Jan 1, 2006

By Denholm W. T

The SMELT-2 package is specifically designed to perform on-line simulation of smelting processes. This paper describes its application to the simulation of continuous matte converting followed by reductive slag cleaning in SIROSMELT reactors. The matte, containing 50% Cu is assumed to be produced at the rate of 20 tph in a continuous matte smelting furnace such as an Outokumpu Flash Furnace, Noranda or Sirosmelt reactor. Two alternative fluxing strategies in the converter are examined, namely: ò conventional silicate slags, ò calcium ferrite slags, and the flow sheets associated with each strategy are discussed. A regular solution model is used in the present study to account for the variations in the activity coefficients of components in the slag.

Jan 1, 1989

By T Goswami, G Brent

In response to concerns over greenhouse gas (GHG) emissions and an imposed cost in some jurisdictions, many mining houses have committed to reductions in their GHG emissions. However, mining faces particular challenges in this regard, notably from increasing strip ratios. Furthermore, mining activity around the globe is not expected to decrease in the foreseeable future as mineral demand is driven up by the aspirations of rapidly-developing nations. Novel approaches will thus be needed if any reductions in energy consumption and GHG emissions of mining are to be achieved.This paper presents some directions for the mining industry to pursue in reducing their GHG impact, focusing on the difference that blasting methods can make. Importantly, large productivity and revenue gains can also be realised from these methods.Key directions considered here include optimising resource recovery and quality through:reducing coal damage and loss from throw blastingrecovering thin coal seamsproducing cleaner coal with less dilution.Blast design can also significantly improve the productivity of surface mining by:improving muck pile shapingimproving mine equipment productivityimproving scheduling.We show how a particular blasting method can eliminate coal loss in overburden throw blasting and can lead to large increases in coalmine profitability. Thin seams that were previously wasted can also be recovered with this method. Furthermore, the method has been implemented to combine several separate drill and blast cycles into a single cycle, leading to productivity gains.An illustrative example is given to show how increased coal recovery can increase revenues and more than offset any potential GHG emissions liabilities. Overall mine energy and GHG emission intensities can thus be reduced.CITATION:Goswami, T and Brent, G, 2015. Blasting approaches to increase mine productivity and reduce greenhouse gas emissions in surface coal mining, in Proceedings 11th International Symposium on Rock Fragmentation by Blasting, pp 635–644 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 24, 2015

This paper reviews recent advances in geophysical tools for mineral exploration. It does not restrict comments to methods which are usually applied in the tectonically active belt surrounding the Pacific Ocean, where geophysics has often been secondary to geological prospecting (eg visual recognition of surface alteration) and drainage geochemistry; but considers geophysical tools applied to a broader range of both deposits and terrains. For example, the subdued topography of much of Australia and Canada has lent itself to the development and extensive application of airborne methods. In Australia, deep weathering and transported cover have necessitated the use of geophysical methods to not only detect anomalies but also to map and understand basement geology. This same weathering has caused specific problems, particularly for electrical methods, but recent improvements in instrumentation, processing and visualisation have largely overcome them. A great deal of geophysical data is now being collected by government agencies, often from the air, and made readily available to explorers. In several Australian states it is free, often on line, and in almost all cases it is relatively cheap. These regional data sets are often augmented by more detailed local surveys in areas of interest. Although æanomaliesÆ are still important much more is being done with the high quality data which is now available The results are used for geological mapping, to test geological hypotheses, to construct detailed 3D geological models and, together with fluid flow modeling, to predict optimum locations for mineralisation.

Jan 1, 2004

The Que River Mine is a small under-ground base and precious metals mine located in north-western Tasmania. Current production rate is 220,000 tonne/yr from reserves grading 0.37% copper, 7.1% lead, 12.2% zinc, 195 g/tonne silver and 3.4 g/tonne gold.Ore is mined using longhole open stoping techniques. Cemented fill is placed in primary stopes to permit safe and efficient recovery of pillars. The operation is trackless with both shaft and decline access from surface.Ore produced is trucked 40 km by road to be sold to the Electrolytic Zinc Company of Australia Ltd., for treatment at their Rosebery concentrator.

Jan 1, 1985

By S B. Sheldon

Wall control programs tend to be expensive and consequently require financial justification in order to get approval for the expenditure required. The difficulty faced in getting this approval is that the outcomes of a wall control program are all subjective – a prediction of wall stability and what the wall would do with and without the program. Determining a net benefit is difficult and often justification of a wall control program is based on gut feel, history and a good selling campaign to build a perceived need.This paper outlines a simple solution to this problem by applying risk management principles to wall control. The process is to identify the risks to the operation associated with current wall conditions and then build a risk database for different potential scenarios for each risk. A financial impact is assigned to each risk scenario to complete the database. A consequence and likelihood analysis of each of the risks is undertaken and a formula applied to estimate the projected cost to the operation for the current risk profile. The process is repeated for the predicted wall control conditions after the application of the wall control program. A successful wall control program will reduce an operation’s risk profile so this projected cost should be less than that calculated for the current scenario. The difference in the two projected costs is the value that the wall control program has produced and this value can be used to determine the net financial benefit potential for the wall control program.This methodology can also be applied to other improvement programs that focus on risk reduction such as vibration control.

Nov 8, 2011

In the traditional analysis of airblast and ground vibration, it is often assumed that only that single blasthole possessing the maximum charge weight (or a small group of blastholes initiating in an 8 ms time window) will determine the peak level of airblast or vibration. This assumption is examined in light of waveform superposition, and is found to be unsound. More rational alternatives are given in which all blastholes (or a selected number of blastholes) are assumed to contribute, in varying degrees, to the peak vibration or airblast. However, it is also shown that there is generally no precise solution for determining the number of blastholes that contribute to the peak level. The technique of waveform superposition does not rely on evaluating which group of holes produced the peak, and so is the preferred method for prediction. Some numerical and analytical models of wave superposition for airblast and vibration are also reviewed and recent approaches discussed. The directionality of airblast, in particular, is analysed in detail using a new superposition model that is constructed of two submodels, onenumerical, the other analytical. This model shows that the airblast from even a single blasthole can be highly directional. The detail of this model is commercially sensitive and will not be discussed. Only results and a broad discussion will be given.

Jan 1, 2004