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By Hirst H, McAfee J

In December, 1934, Mr. R. Sussex extruded some silver lead alloy wires in order to determine their electrical resistance. After the determinations had been made each wire was threaded through a sheet of paper on which the composition was noted. The wires were placed in a more or less tangled mass in a cupboard in March, 1935. It was not until September, 1936, that they were re-examined, and it was found that during the eighteen months some of the wires had corroded very badly. It was not until March, 1938, that the wires and the corrosion product could be examined, and the results are given in the following pages. It is of interest to note that none of the numerous silver-lead alloys which have been used for creep testing in the rolled condition have shown a similarphenomenon. The silver content of the rolled alloys was, however, never greater than 0·1%.GENERAL APPEARANCE OF THE CORRODED WIRESThe wires contained 0·1% up to 28% of silver, and a general examination showed that the corrosion had progressed farther the higher the silver content. This can be seen in Fig. 1, which is a photograph taken in March, 1938, of typical wires at a 'magnification of five diameters. In the early stages the corrosive action seems to be localized and an area of grey, yellow or reddish brown powder appears. The grey corrosion product is by far the one, most frequently seen, and in the later stages the only one visible from the outside. The reddish coloured product is sometimes seen on the residual metallic centre when the grey product has been broken away.The irregular extent of the corrosion can be seen in Fig. 2, which shows (natural size) the 5% silver wire after standing two years. The corroded parts are seen to be much thicker than the slightly corroded ones. The corrosion product, being non-metallic, is quite brittle and falls away from the wire on bending. In the extreme case the' whole wire has corroded and presents a cracked appearance as seen in Fig. 1 (281% silver).An approximate idea of the extent of corrosion can be obtained by measuring the diameter of the corroded part. The original diameter in each case was ·050/ ·051 inches and in Table I are given the diameters of the corroded parts after Hand 3 years respectively. When the diameter...

Jan 1, 1941

By D Yates, C C(te, M Rodgerson

This paper describes how specific technical and regulatory issues related to water management have been dealt with at Anglo AmericanÆs Moranbah North mine. It compiles the collection of practices that have been implemented since 2010 in response to climatic and regulatory drivers. Operations must continually evolve as they face emerging challenges. For Moranbah North mine, one key challenge is the integration of its mine water system to that of its new neighbouring mine Grosvenor. The commitment to ongoing improvement has placed the mine in a good position to face this particular challenge.Water-related risks were identified then mitigated by technical practices, which have focused on improving underground water management, increasing storage capacity and installing a state-of-the-art control system for the management of water on-site and for releases of mine-affected water to the environment. These practices are supported by an extensive monitoring and audit system that provides confidence in the reporting of activities and outcomes. The strategy has helped the operation with its continual attempts to minimise reliance on high-quality third-party water and improving its interaction with the surrounding environment and community.CITATION:Rodgerson, M, Yates, D and C(te, C, 2013. Leading practice water management at Moranbah North Underground Mine, in Proceedings Water in Mining 2013 , pp 301-310 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 26, 2013

The objectives of this paper are to introduce the International Code of Practice for Risk Management of Tunnel Works and to illustrate to the participants of the 13th Australian Tunnelling Conference 2008 how the insurance industry would like to see the future of tunnelling insurance. Tunnel construction projects have suffered large losses over the years as can be shown from various loss examples and a list of major losses over the last ten to 15 years. Since the consequences of insufficient risk control had to be borne by insurers, these losses have heightened the need for improved and adequate risk management measures so that insurance cover can also in future be provided for this hazardous industry sector. Observations are made as regards to the prevailing trends in the tunnelling industry, which contribute to the increase in risk and have adverse consequences for the insurance industry. To find a solution to this dilemma various options are examined with the suggestion of a Joint Code of Practice for Tunnel Works being agreed by all stakeholders, which would become essential to secure future insurance protection for such project work. The British Tunnelling Society has, in cooperation with the Association of British Insurers, developed such a code, which the International Tunnelling Insurance Group has now adopted for global application in international markets. The content and concept of this code is explained to establish a framework of understanding and applied risk management principles for tunnel works. This initiative also has the support of the International Tunnelling Association and the International Association of Engineering Insurers. The international application of this new code is demonstrated with a few examples, which also includes Australian projects. It is the aim that this code forms in future part of the insurance coverage for tunnelling works in order to put this risk segment on a more sustainable basis. For that reason the insurance industry is seeking the support of all stakeholders of tunnelling construction projects.

Jan 1, 2008

llmenite from the west coast of the South Island, New Zealand, shows exceptional reactivity with dilute sulphuric acid. Based on this, a cyclic acid leaching process was developed for the production of material with more than 98% TiO2, suitable as a substitute for mineral rutile.Titanium hydrate is recovered from the leach liquor by hydrolysis, and is calcined to give the required product.Ferrous sulphate is crystallized from the hydrolysis filtrates, which after concentration are returned to the leaching stage. The need for further evaluation of the process on a larger scale is indicated.

Jan 1, 1973

A poured slag bottom is universally considered to be the best type bottom for copper reverberatory furnaces. It is impervious to molten matte and ready for immediate use.When molten slag is not available a monolithic silica bottom is the second best; such a bottom has been adopted at Mount Morgan. The laying of such a bottom is a phase of smelter construction on which so little real information is obtainable that a day to day record of such a recent example as Mt. Morgan's 19 ft. x 40 ft. reverberatory may be of interest. The work was done under the direction of E. Pesout, of McGill Smelter, Nevada. PREPARATION OF THE BOTTOMA 26-inch bed of closely packed firebrick with sand-filled joints was first laid on the clean, dry concrete foundation and the whole dried out by small fires maintained on the brickwork for several days. Ninety tons of silica sand was meantime thoroughly dried and stored in clean drums. This sand must be low in FeO and CaO and contain about 95% Si02 , the majority of which should be present as tridymite. This form of silica shows little expansion at high temperatures, for it is the form stable between 870° and 1470° C.Its presence thus minimises disturbance to the bottom by the expansion which, during the burning in, would mark the transformation of quartz to tridymite.

Jan 1, 1939

It is now five years since the advent of gold-dredging in Australia, the State of New South 'Wales, or colony as it was, being the first to adopt this system of gold-getting, and the credit of its introduction rests with a member of this Institute, Mr. Chas. L. Garland, who commenced operations with his dredge on the Macquarie River, at the junction of the Muckerawa Creek, in the month of March, 1899. In this paper it is intended to give a brief history of the career of the gold dredge in New South Wales from that relate up to the present time, during which period the writer, as a designer of many dredges, has been constantly in touch with the subject in all its phases in this State; the points noted are the outcome of his observation in a general way, and to a great extent are apart from the design proper.Gold-dredging in New South Wales up to the present moment can hardly be designated an unqualified success...

Jan 1, 1904

By K Kermanshahi, J Allen, MáHilden, M Powell, S Mwansa, S Palaniandy

Approximately 420 VertiMills« (VTM) have been installed in mineral processing plants throughout the world, mainly in tertiary and regrind applications plus several in secondary grinding duties. Metso Minerals (Metso) and the Julius Kruttschnitt Mineral Research Centre (JKMRC) have commenced a collaborative research program in stirred milling technology focusing on mill performance evaluation, model development and scale-up methodology. Industrial VTMs are being assessed using a consistent circuit survey protocol to evaluate the performance of VTMs in comminution circuits. The survey protocol includes identification and modification of sampling points, appropriate sampling techniques, controlled circuit stability, and best sample analysis practice. To initiate this study four surveys were conducted in the tertiary grind circuit in the Ridgeway Concentrator at Cadia Valley Operations (CVO) that treats the full circuit throughput of over 800 tph. To deal with the high feed rate the site opted to lead the world in the application of stirred mills by installing the largest VTM, at 3000 hp (2240kW) in closed circuit with a hydrocyclone bank. On average the operating work index was 13.1 kWh/t, reducing the secondary grind product by 40 - 50 ¦m to the required 80 to 100 ¦m range. The applied specific energy is relatively low in these low-intensity stirred media mills at 2.7 kWh/t. The size specific energy was 24.2 kWh/t-75¦m. Based on the survey and operating data the VTM circuit is performing well in this high-throughput tertiary grinding application, producing a sub-100¦m product.CITATION:Palaniandy, S, Powell, M, Hilden, M, Kermanshahi, K, Allen, J and Mwansa, S, 2013. VertiMill« - Development of circuit survey and performance evaluation protocols, in Proceedings MetPlant 2013 , pp 162-175 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 15, 2013

The dynamic response of holdup in a continuous ball mill has a direct influence on the rate function of grinding kinetics and is useful for predicting the variation in the size distribution of the finished product with time.A model was proposed involving three interacting, completely mixed chambers and this was used to derive analytically the transfer function of the system, from which the unit impulse response and the unit step response of the holdup with feed rate were obtained.The theory was confirmed by comparing it with previously published experimental data for the residence time distribution of particles passing through a ball mill.

Jan 1, 1975

The arbitration process, and its drawbacks: Any time you have taken part in a major arbitration, and the decision, when it comes out, is not what you thought it would be (which is bound to happen occasionally), the normal reaction is something like æthe arbitrator got it wrongÆ. Again normally, when you have cooled down a bit, and you look in more detail at the decision, you can see, probably, æwhere he got it wrongÆ (it is only male arbitrators who get it wrong). Ultimately, when reason prevails, and you reflect on it, you can see that what really happened was that you failed to get the arbitrator to see what you were getting at in the point or points concerned. This is doubly frustrating, when you consider what has been involved in getting to this point, having been through it all at least twice. An alternative to this process may be the dispute review board.

Jan 1, 1999

By H He, S Tillotson, S Gibbons, J Shaw

"Water evaluation and management during mining exploration and operational activities is crucial to minimise direct and indirect operational costs and also in delivering mining companies the less tangible, but equally critical æsocial license to operateÆ.Mining is often carried out in remote areas, where the mine will be the most significant and visible user and potential contaminator of water resources. However, wherever it is located, a mineÆs use of water must be managed in the context of two key aspects: the natural parameters of the resource; its abundance, its temporal variability and its chemical and physical qualities other users of the resource; other industry, potable water supplies and ecological requirements of local habitats.It is vital, therefore, that this commodity be assessed to an appropriate degree to help establish the viability and sustainability of the mining project and identify those other stakeholders who may be affected by the mining companyÆs activities.This paper outlines the background to catchment-scale model and presents a cost-effective method for establishing a robust water baseline and creating a water management tool to be used throughout whole mine life cycle. Numerical surface and groundwater modelling has long been established as a tool for reactive management, following collation of yearsÆ worth of data. Now, using freely available remote sensing data and desk study information, analytical modelling techniques can be used cost effectively to provide mining companies with a tool for assessing and managing their water consumption requirements at the outset of a new project venture.The paper includes a selection of case studies to show how the use of analytical modelling can be used as an evaluation and management tool to help not only assess general watershed parameters and water availability but also how it can be used as a tool to provide decision makers with the information to complete æwhat if?Æ scenario assessments. The case studies illustrate how a combined hydrological and hydrogeological model may enable identification of sensitive receptors, water competition issues and contaminant migration pathways to help manage the risks associated with water consumption and its potential contamination from exploration and operational activities.CITATION:Gibbons, S, Shaw, J, Tillotson, S and He, H, 2013. Catchment-scale water modelling and its role in risk management during mining exploration, in Proceedings Water in Mining 2013, pp 175-188 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Nov 26, 2013

By G F. Brent, R D. Hart, C Carr, F Felipe

In underground operations, where copper, gold and other metals are mined, the open stopes can be backfilled to stabilise surrounding workings. This fill may consist of tailings from the processing plant, usually mixed in slurry with a binder such as ordinary Portland cement (OPC). Extensive use is made of tailings in OPC-based paste fill, cemented hydraulic fill or cemented aggregate fill applications. A large underground operation such as the Mount Isa Copper Operations utilises over three million tonnes of backfill per annum. Geopolymers offer an alternative binder system with significant environmental advantages over OPC. Geopolymers are formed by polymerisation of aluminosilicate precursors in an alkali medium. They generally form products with superior strength characteristics as well as increased resistance to degradation, particularly under acidic conditions. Geopolymers have lower carbon dioxide emissions associated with their manufacture than OPC and offer further potential advantages, such as the use of less water and more waste materials. The application of geopolymers to weak material systems such as mine backfill has not been widely studied. The aims of this investigation were to assess a range of Australian mine process wastes for suitability as feedstocks in fill applications and to construct a range of these cemented minefills using a geopolymer binder. The physicochemical characteristics of each of the wastes are discussed as well as how they ultimately contribute to the geopolymer cement through direct replacement of OPC. Results from various tests on the candidate geopolymer fill formulations are presented and the potential of such fill formulations to replace OPC in mine backfilling operations is highlighted. The significant potential contribution to reducing greenhouse gas emissions from the use of geopolymer-based mine backfill across the industry is noted. Conversion of a single mine, such as the Mount Isa Copper Operations, is shown to match the reductions in greenhouse gas emissions due to the current total installed photovoltaic electricity generation capacity of Australia. A similar examination of wastes available to a gold mine could potentially yield new binders for use in established or new fill systems.

Jan 1, 2007

Laboratory and pilot plant investigations aimed at recovering gold as an electroformed sheet from gold eluate and gold leach solutions and slurries are discussed. The electrolytic test work undertaken makes use of high surface area electrodes, in particular nickel-chrome foam metal and activated carbon, as well as ultrasonic activation. The process involves initial gold electrowinning (or adsorption onto activated carbon), followed by gold electroforming, both operations being carried out in a specially designed reactor. The pilot plant work mainly involved the treatment of gold eluate solutions from a South African gold mine containing 600 mg/L Au: some tests were also carried out on gold leach solutions containing 6 mg/L Au. In the former case, very efficient gold extraction onto the foam metal cathode was achieved using electrolysis combined with ultrasonics. The polarity of the gold-loaded cathode was then reversed and, using a replacement electroforming electrolyte, gold was electroformed in situ onto a polished stainless steel mandrel, from which gold sheet was subsequently removed. In the case of leach solutions, pilot plant tests confirmed laboratory findings that in the presence of calcium (resulting from the lime used during the leaching process) no electrowinning is possible. Laboratory tests showed, however, that when sodium (in the form of NaOH or Na2CO3) is added to the leach solution, electrowinning can indeed take place. Alternatively, if a special activated carbon electrode is used (without current), the non-ionic calcium aurocyanide complex is adsorbed on to the activated carbon. When a gold-loaded activated carbon electrode is made anodic in a gold electroforming electrolyte, gold electroforming is feasible, although the efficiency is very low. The possible implications of these developments for gold recovery are further examined.

Jan 1, 2007

By R L. Brathwaite

Previous paleodepth studies of low-sulphidation epithermal Au-Ag deposits in the Coromandel region indicated that economic gold mineralisation was generally limited to a depth zone of 100 to 400 m below the paleosurface. Notable exceptions are Waihi and Karangahake, where ore shoots extended to depths of about 750 and 900 m respectively. These studies have been extended to include Au-Ag, Hg and Sb epithermal deposits in Northland and the Taupo Volcanic Zone. The main paleodepth indicators used are sinters and hydrothermal eruption breccias, hydrostatic pressures estimated from homogenisation temperature measurements on fluid inclusions, and hydrothermal alteration mineralogy. Also the occurrence of lacustrine and fluvial sediments is an indicator of the paleosurface at their time of deposition. Selected paleodepth indicators have been compiled and mapped in a GIS environment. The distribution of the paleodepth indicators are used to define regional variations in erosion levels, and may be useful in predicting areas most likely to preserve the optimum depth window for gold deposition. In the Coromandel region, regional stream sediment geochemical data has also been used to broadly define erosion levels in terms of Zn, Pb and Cu representing deeper base metal zone mineralization and Hg and Sb representing shallower cinnabar and stibnite mineralisation. The interpretations of paleodepth and potential reconstruction of paleosurfaces are limited by the paucity of data, poor age control on the paleodepth indicators and the epithermal deposits, and marked relief and complex stratigraphy in the volcanic terranes.

Jan 1, 2003

By Nicol S. K, Firth B. A

Much effort over the years has been put into the task of devising parameters which adequately assess the metallurgical efficiency of mineral processmg operations. One of the most celebrated of these methods is the so-called "release analysis" method of Dell (1953) which tackles the problem in terms of an ideal separation based on sequential retreatment of the products and tails of a given separation process. Other methods have also been proposed by Jowett (1969) and by Holland-Batt (1971 ). While these procedures are appropriate to multi-component mineralogical separations, they perhaps do not lend themselves to the much more simple two-component systems encountered in coal preparation. However Dell (1962) has suggested the use of the term "specific profit" in coal preparation in order to assess the conditions that give maximum economic gain. Specific profit is derived from data relating washability data with equipment performance and operating costs. The procedure for using this criterion for comparative purposes is therefore somewhat cumbersome as is illustrated by its lack of widespread acceptance in the coal preparation literature.In contrast, an efficiency index, E, which is both simple and meaningful from the utilitarian standpoint, has been proposed by Tsiperovich and Evtushenko (1959a). This index is defined by the relation, E = Yield X Tailings ash/Product ash (1) A cursory examination of this parameter shows that it increases when yield and tailings ash increase and also when product ash decreases; all of which are desirable circumstances in coal cleaning operations.Further, it will be shown that slight variations in operation are amplified by this function which makes comparisons simple and direct.While the use of this index is quite extensive in the Soviet coal preparation literature for some reason it has not received wide usag~ elsewhere.

Jan 1, 1978

By P Holtham, K Duffy, W Valery, A Jankovic

As high grade ore deposits are depleted, remaining deposits are generally lower grade and require the mining and processing of larger volumes of material per tonne of product. These low grade deposits often contain a large proportion of barren gangue. Pre-concentration aims to remove this barren material at as coarse a particle size and as early in the process as possible. This has the potential to reduce processing costs, energy and water consumption, increase production and improve resource utilisation.There are several technologies that may be applicable for pre-concentration. This paper looks specifically at bulk ore sorting and how it can be integrated into existing or new mining operations to improve profitability and reduce environmental impact. Bulk ore sorting is the separation of barren gangue from a fully loaded conveyor belt based on the grade as measured or inferred from a sensor measurement. It may be used to increase the production rate through fixed plant capacity, or increase production and/or reduce the size of downstream processing equipment in Greenfield operations. Marginal ore sub-economic reserves may be upgraded, making them economic to treat and improving the resource utilization. Bulk ore sorting could also be used to reduce dilution and ore loss in mining operations by improving grade control, or to separate ore types to treat via different process routes. Various applications of bulk ore sorting are discussed in this paper, and three cases are investigated considering the impact on the overall operation and the associated benefits.CITATION:Duffy, K, Valery, W, Jankovic, A and Holtham, P, 2015. Integrating bulk ore sorting into a mining operation to maximise profitability, in Proceedings MetPlant 2015, pp 273–287 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 7, 2015

The early records of hydraulic elevator and bucket ladder dredging in the Clutha River system indicate the success of that era of alluvial gold mining in Otago. This paper briefly reviews parameters of depth and gold grade for alluvial operations. The application of dragline technology is proposed for the economic removal of overburden to target the deeper alluvial gold areas which remain in the Otago region.

Jan 1, 1992

Mining Systems and Technology - Synchronous Electric Drives for Grinding Mills

Sep 13, 2010

The richer, shallower sapphire deposits of the New England region of New South Wales which have been traditionally mined by the small miner are almost worked out. Larger mining and exploration companies are replacing the small miner and exploring for the primary source of sapphire or lower grade alluvial deposits.   Although in recent years considerable exploration effort has been directed at locating the primary source, commercial exploitation of sapphire is still restricted to the working of alluvial deposits.   If the Australian industry is to compete against Asian and African producers it must use cost effective exploration, mining) processing and marketing methods.

Jan 1, 1990

By F Gervilla, J Laukkanen

The Kevitsa Ni-Cu-PGE deposit in northern Finland contains disseminated Ni-Cu-Fe sulfides, and platinum-group minerals (PGM). The 2.05 Ga-old ultramafic-mafic, funnel-shaped intrusion is intruded into a volcanosedimentary sequence. The major ore minerals are pentlandite, hexagonal pyrrhotite and troilite, chalcopyrite, cubanite and magnetite. Pyrrhotite and pentlandite are replaced by pyrite, millerite and heazelwoodite near the surface. Nickeline, maucherite and gersdorffite are locally abundant. A pilot feed sample of 575 t of the main orebody at Kevitsa had average grades 0.45 per cent Cu, 0.36 per cent Ni, 1.85 per cent S, 0.33 g/t Pt, 0.34 g/t Pd and 0.19 g/t Au, with 0.06 per cent Ni contained in silicates. Higher PGE contents (2 - 27 ppm) occur in a vertical N-S bearing fault zone. The PGM (n = 6022) mapped in specimens of four drill holes, consist of sperrylite 34.7 vol per cent, braggite and cooperite 18.5 vol per cent, geversite 17.2 vol per cent, moncheite 14.7 vol per cent, merenskyite 6.7 vol per cent, stibiopalladinite 4.3 vol per cent, sobolevskite 1.9 vol per cent, kotulskite 0.7 vol per cent, michenerite 0.6 vol per cent, keithconnite 0.3 vol per cent, rustenburgite 0.1 vol per cent, melonite, froodite, laurite together 0.1 vol per cent and only a few grains of sudburyite, irarsite, vysotskite, erlichmanite, hongishiite, Pt-Cu-alloys, three undefined Pd and Pt bismuthinides and antimonides and a Pd-Cu-bearing Pt-Ni-Fe sulfide. Most of the PGM, 62 per cent, are enclosed in silicates, 29 per cent are intergrown with sulfides, nine per cent are within sulfides. A portion of the PGE occurs in solid solution in sulfides and arsenides. Ni flotation yielded smelter-grade concentrates averaging 12.2 per cent nickel. The average sulfide Ni recovery is 78.8 per cent and total Ni recovery is 65.3 per cent. The average Pt is 10 g/t at 55 per cent recovery, Pd 7.7 g/t at 38 per cent recovery and Au 1.6 g/t at ten per cent recovery. Cu concentrates averaged 28.6 per cent Cu with an average recovery of 78 per cent and further 11 per cent in the nickel concentrate. The copper concentrate contains 4 g/t Au at 24 per cent recovery. The average Fe to MgO ratio 17.9 is well above the targeted minimum of 4.5. The pilot results show that the previously used TETA (triethylene tetramine) can be avoided and good quality, smelter-grade concentrates can be produced from the Kevitsa ore at economical levels of recovery by conventional flotation methods.

Jan 1, 2008

By T Howard, J E. Everett

BHP Billiton Iron Ore has replaced the conventional product quality control batch system at Yandi with the Continuous Stockpile Management System (CSMS) centred on a fully integrated æcontinuousÆ stockpile flow from mine planning through to port reclaiming. With increasing tonnage rates, tight operational coupling associated with the original batch system caused production interference which achieved final product quality but in an increasingly inefficient manner. This innovative approach places a continuous emphasis on product grade control and takes full advantage of the most appropriate blending opportunities of the mine to port planning and production system in order to reduce the natural variability of the ore to an acceptable level. To implement continuous control, product variability throughout the process is classified as long-term and short-term. Long-term variability is best controlled through the monthly mine planning process which determines monthly grade targets while taking into account optimum mine development. In the past, attempts to control long-term variability were made at the port through selective stockpiling and reclaiming which reduced operational efficiency. Short-term variability is best controlled around the monthly mine grade target by daily mine production scheduling, train sequencing, stacking into port stockpiles of appropriate size and systematical stockpile reclaiming. The implementation of the new system has maintained customer accepted shipping grades and has provided significant improvements in resource utilisation, operational efficiencies and costs.

Jan 1, 2002