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By T Takayama, M Matsumura, Y Konno

The steel industry accounts for 14 per cent of domestic CO2 emission in Japan, and iron ore sintering process occupies approximately three per cent. While Fe content in iron ore is gradually decreasing. So, the amount of concentrated iron ore will increase in the future. We have focused on magnetite concentrate as sinter materials, because magnetite can be easily magnetically beneficiated. However, it is experienced that blending magnetite ore resulted in increasing Fe2+ component in the iron ore sinter, which results in lower reducibility. Therefore, promoting the oxidation reaction from magnetite to hematite in the sintering process is countermeasure for reducibility. In addition, this promotion causes increase of oxidation heat, which results in high sinter strength and reduction of coke. It is assumed that the oxidation of magnetite tends to proceed in the cooling stage, where the oxygen partial pressure increases after the combustion of fine coke, but there are few studies on the effect of heat profile, especially cooling path. The purpose of this study is to design the effective sintering heat profile to promote oxidation of magnetite ore, especially, to investigate the effect of cooling path on sintered ore microstructure. In this study, magnetite ore and magnetite reagent were respectively mixed with calcium carbonate reagent. Then the mixed samples were pressed and formed into cylindrical tablets for simulating the adhesive layer of a pseudo particle. Through examining influence of maximum temperature and cooking rate on oxidation reaction of magnetite, we confirmed that the oxidation reaction of magnetite is suppressed at above 1350°C for maximum temperature, compared to 1300°C. Sintering at lower temperature is desirable because hematite is stable below 1350°C in the Fe2O3-CaO phase diagram.

Nov 8, 2021

By K G. Harding, A Osman, A du Toit, F K. Crundwell, C M. Sheridan

"Most mineral processing plants have high water requirements for their operational activities. These plants often function in an environment where water is becoming increasingly scarce. An increase in population will result in an increased demand for water; potentially beyond the limits of supply and hence increased competition for the resource. In South Africa, Gauteng and the North West Province are likely to be first to experience a shortage of potable water.In order to avoid a water crisis, efforts must be made to ensure the effective management of water resources and the minimisation of wastage. Hydrometallurgical plants have a responsibility to society and the environment to operate in a sustainable manner - reducing their water requirements will certainly assist towards developing a sustainable footprint.This project improved site water management by: surveying the potable water and storm water systems with a view to assessing the current water accountability and determining methods to improve accountability determining the monthly and yearly water balance for the site and presenting it in the form of a water balance sheet drawing up an input-output and operational model in accordance with the Minerals Council of AustraliaÆs æWater Accounting Framework for the Minerals IndustryÆ calculating the water footprints of the refinery identifying ways in which water savings could be achieved recommending management systems for water usage monitoring.This work was conducted in South Africa. The value of this work lies in that it established the current baseline and assisted in developing management systems that can reduce water usage sustainably.CITATION:Osman, A, Crundwell, F K, Harding, K G, Sheridan, C M and du Toit, A, 2013. Water accountability and efficiency at a base metals refinery, in Proceedings Water in Mining 2013, pp 277-286 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Nov 26, 2013

By M Freeman

In a visionary move, the Planning Commission and the Ministry for Planning in Western Australia decided to invite three extractive industry quarry operators, local government councils and State Government departments (Minerals and Energy, Environmental Protection, South West Development Commission) to contribute jointly to study the impacts of blasting on areas surrounding basalt quarries in the Bunbury region of WA. The results of the study will help establish suitable zones in which land uses will be controlled under a proposed Greater Bunbury Region Scheme and in local planning schemes, to minimise conflicting land uses. This paper discusses the blasting program conducted to characterise and model the vibration behaviour of the rock under the effects of pilot-scale and full production blasting, and highlights some of the benefits achievable when a thorough blasting impact program and assessment is conducted at the planning stages of land-use establishment.

Jan 1, 2001

By D van der Spuy, T Kewe, K Keet, N Moffatt, A P. Paine

The Blue Cube Mineral Quantifier Inline (MQi) instrument uses reflective light spectroscopy and advanced chemometrics to measure mineral or elemental grade in a slurry line. The instrument takes scans continuously and sends data to the DCS or PLC control system.Controlling the concentrate grades on a flotation circuit using low-frequency measurements can be difficult or impossible, especially on a circuit that has fast process dynamics. The ability to rapidly measure changes in grade makes it possible to continuously manipulate the flotation circuit operating parameters such as level, airflow and reagent addition set points to optimise plant performance throughout changing conditions.At Barrick’s Porgera mine in Papua New Guinea the flotation circuit feeds sulfide concentrate slurry to a pressure oxidation circuit. In order to maximise gold recovery in the flotation plant and increase performance in the pressure oxidation process, there is a stringent requirement to maintain a stable and optimum sulfur grade. The gold recovery is directly affected by the sulfur grade as most of the gold at Porgera is submicroscopic and associated with the sulfides (disseminated?pyrite).Porgera considered various points in the flotation circuit for installing Blue Cube analysers. This was part of a plant upgrade to improve process control. Due to budget constraints it was decided to initially only install one Blue Cube analyser on the final concentrate stream of the flotation circuit measuring sulfur, sulfides, gold and particle size. The accuracy of the calibration results were very good for all these properties. The sulfur measurement was used in an advanced control system to optimise and stabilise the grade. The paper presents comparisons between mineral grades determined using analytical laboratory analyses and the online Blue Cube data measurements, as well as benefits achieved by measuring grade online. An analysis of the plant data before and after installation of the Blue Cube and the advanced process control upgrade indicate that the flotation recovery increased by 1.98 per cent.CITATION:Kewe, T, Moffatt, N, Strobos, P, van der Spuy, D, Paine, A P and Keet, K, 2014. Evaluation of the Blue Cube MQi slurry analyser for application in an advanced control system for the optimisation of a gold sulfide flotation circuit, in Proceedings 12th AusIMM Mill Operators’ Conference 2014 , pp 357–362 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 1, 2014

By D Brennan, S Jahanshahi, F Jorgensen

The recycling of metals can be very efficient with careful selection of the material being recycled. The volume and value of recycled ferrous scrap has lead to the development of steel mini-mills which in turn has caused major changes to the steel industry. Although much smaller in volume, the recycling of lead is very significant, with battery recycling accounting for 60 per cent of annual world production. This high recycle rate can be attributed to the relatively uniform feed-stock from lead-acid batteries. The recycling rates of other metals are considerably lower due to longer use and non-uniform product usage that results in less efficient segregation and processing. It is clear that recycling of metals has advantages but its sustainability is uncertain. What are the key benefits of metal recycling? Metal recycling has technical, economic and social impacts. It is relatively easy to assess recycling on a technical basis, but the social impacts are harder to evaluate. This paper examines some of the issues in evaluating the sustainability of metal recycling.

Jan 1, 2002

By Frfas C, Parrilla F, Regife J. M

A project to develop the Zinclor electrowinning cell to industrial size is being conducted at Tecnicas Reunidas' Research and Development Centre with Spanish Ministry of Industry support. This project is the final development phase in zinc electrowinning from a chloride medium, the highest ZINCLOR technology development level prior to the industrial scale having already been reached. This paper covers the outcomes and conclusions corresponding to project Phases 1 and 2: Process Selection and Industrial Electrodes (1 m2) design and testing. The most important achievement reached is that there are presently available industrial cathodes and anodes for the Zinclor diaphragm cell giving an optimum response for zinc chloride electrowinning with unit power consumptions between 2.65 and 2.80 kWh/kg Zn, depending on the diaphragm type. Standard electrolysis conditions were: 65 g/L ZnC12, 3 g/L HCl, 160 g/L NaCl, 350 g/t Zn arabig gum as the addition agent, 36¦C, and 400 A/m2 current density. The selection of electrodes, diaphragms and organic additives for the industrial scale tests have been optimised. The complete electrowinning cell development, with electrodes having an active area of I m2, enables the evaluation of Zinclor Technology application for industrial processes.

Jan 1, 1993

By G R. Dickens, N H. S Oliver

The genesis of giant hematite ore deposits in northwest Australia remains a contentious topic in part because key evidence supporting a unifying genetic model has not been found at all deposits. In particular, several papers have proposed that carbonate replacement of silica layers in banded iron formation (BIF) by basinal brines is a requisite first step toward ore formation. This initial hypogene stage is inferred primarily from the presence of silica-poor, carbonate-rich rocks found below ore and along normal faults at the Mount Tom Price and Giles Mini deposits. However, until recently such rocks have not been identified at the largest martite-microplaty hematite deposit, Mount Whaleback. In this study, samples of the upper Mount McRae Shale are examined for their chemistry, mineralogy and petrography. These samples were collected from several key locations, including within drill core that penetrates an area immediately beneath ore along the Mount Whaleback fault at Mount Whaleback. Compared to unaltered black shale from Wittenoom Gorge in the north and altered black and red shale at Mount Whaleback, reddish-green shale below the Mount Whaleback deposit is significantly enriched in MgO and CaO and depleted in SiO2. Samples of this shale consist predominantly of fine- to medium-grained dolomite and ankerite cut by chlorite and carbonate veins, contrasting with unaltered equivalents that contain mostly stilpnomelane, K-feldspar and relatively minor carbonates. This alteration is distinct from assemblages developed during regional metamorphism, and possibly represents hydrothermal alteration after metamorphism. Although several questions remain unanswered, these results support models that invoke an early hypogene stage during the formation of the martite-microplaty hematite deposits in the Hamersley Province.

Jan 1, 2005

By Zhao JC, Chen JJJ

A water/air model was developed to simulate the refining of molten aluminium by the sparging of gas through an impeller. The optimal gas dispersion depends upon selecting the correct gas rate for a given impeller rotational velocity to prevent flooding or channelling. Correlations illustrating this flooding behaviour exist for standard impellers within standard tank geometries. However the validity of these correlations has not been substantiated for gas dispersion in high temperature molten metals with non-standard tank and impeller dimensions.

Jan 1, 1995

By S Vink, A Wolhuter

Increasingly the community and government regulators are calling for æzero dischargeÆ mines. Some mining companies are beginning to explore ways in which they can transition to zero discharge. When striving to attain zero discharge, mines must consider multiple variables in addition to water quantity. These include climate regime, water quality and regional water budgets. This paper explores the drivers behind the concept of zero discharge mines and will use examples to illustrate the feasibility of attaining this goal.CITATION:Wolhuter, A and Vink, S, 2013. Is zero discharge a viable or useful concept in mine water management?, in Proceedings Water in Mining 2013 , pp 329-334 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 26, 2013

RME has a vision for a new grinding plant operational paradigm made possible by RMEÆs fast, safe Mill Relining System. Some new liner designs can degrade SAG mill performance because of mill volume loss, ball strike (rotational speed limitations) and other issues often brought about by modifying liner design for maximum life and without a backward glance at metallurgical consequences. Freshly relined mills can initially produce as little as 80 per cent of their maximum throughput, that maximum occurring toward the end of the linersÆ wear life. RMEÆs proposition is to see new liners designed for maximum commercial performance. This may mean more frequent mill relines but with liners designed to yield optimum grinding, metallurgical and profitability performance, not maximum life. In other words, mill liner profiles and mill relining speed become key operational controls and not simply æmust doÆ maintenance functions. Taken to the extreme, liner profiles may prove capable of ætuningÆ a grinding plant for maximum profitability in any metal price environment. Following the introduction, the first section of this paper provides an update describing recent and significant developments within RMEÆs range of Mill Relining Systems, the enabling technologies which permit more frequent liner exchange without incurring additional downtime imposts. The second section outlines RMEÆs vision of a æliner profile versus profitabilityÆ commercial model, developed either intuitively or by using a combination of new measurement and modelling technologies. These new measurement and modelling tools provide the opportunity to predict the optimum commercial outcome, off-line in the virtual world, before committing to an expensive liner experiment.

Jan 1, 2007

By M I. Pownceby, M A. Rhamdhani, D M. Fellicia, S Palanisamy, R Z. Mukhlis

The trend of global copper production has prospectively increased over time. Based on typical mined ore grades, 1 t of copper ore generates approximately 6–10 kg of copper, which requires much energy, usually in the form of metallurgical coke. Copper production using carbon as a fuel and reductant contributes up to 0.3 per cent to global greenhouse gas (GHG) emissions. As a result, research into decarbonisation processes applicable to reducing low-grade copper sulfide ores, copper oxides, electronic wastes and other alternative complex Cu-rich materials has increased. Alternative fuels for reducing carbon-rich emissions in pyrometallurgical copper processing include methane, ammonia, biomass, solar and wind power, but in recent times increased focus has been on hydrogen as a potential fuel and reducing agent for materials such as oxidised copper scrap, Cu2O/CuO and Cu-rich slags/e-wastes. From a thermodynamic perspective, hydrogen exhibits a significantly more negative standard Gibbs free energy (ΔG°) than copper oxide making it a suitable reductant and the exothermic thermal effect from reaction between hydrogen and Cu2O/CuO may be used to control process parameters. These characteristics renders hydrogen an ideal gas for reducing copper oxides and copper-containing slags/e-wastes. This review article assesses previous research on utilising hydrogen for producing and refining copper from primary and secondary feed materials.

Jun 19, 2024

By G J. Lyman, S A. Tittlemier

International regulations exist to manage the concentrations of mycotoxins such as Ochratoxin A (OTA) and deoxynivalenol (DON) in grains and prevent health risks from excessive exposure to OTA and DON. Since the permitted levels of these mycotoxins are very low (existing maximum levels are in the range of 5 ppb for OTA and 1–2 ppm for DON), sampling for these mycotoxins is potentially challenging. Given that the distribution of infected kernels in a consignment of wheat can be highly heterogeneous, detection by current sampling methods is subject to a high variance. Furthermore, the concentrations of mycotoxins in individual kernels vary over orders of magnitude, intensifying the heterogeneity of the bulk grain.In these circumstances, there is much interest in developing sampling plans that inform regulatory agencies, and mitigate the risk of misclassifying grain shipments as compliant or non-compliant within the global grain trade, when correctly implemented.Experimental data from laboratory and field studies describing individual wheat kernel concentrations of OTA and DON were combined with experimental data describing the variance due to the analytical methods used for OTA and DON to calculate the sampling distribution for wheat having various levels of contamination. This work has focused on quantifying the heterogeneity of samples at the point of submission for analysis and within subsequent sample preparation to arrive at total sampling distributions that include uncertainties due to the intrinsic heterogeneity and the analytical uncertainties. The sampling uncertainty that relates to the difference between the primary sample analysis and the true lot analysis is not included as it has been dealt with elsewhere.CITATION:Lyman, G J and Tittlemier, S A, 2017. The sampling characteristics of grains contaminated by mycotoxins, in Proceedings Eighth World Conference on Sampling and Blending , pp 281–294 (The Australasian Institute of Mining and Metallurgy: Melbourne).

May 9, 2017

By T Kon, K Ohno

Slag foaming is an important control item in the process of refining molten steel. This study aims to investigate the factors controlling slag foaming by observing the bubble formation behaviour resulting from the chemical reaction between iron oxide and Fe-C alloy in the slag. In this study, 0.06 g of Fe- C alloy was introduced at the bottom of a BN crucible, and 6.0 g of slag (SiO2:CaO:Fe2O3 = 40:40:30) was placed on top of it. The crucible was positioned inside an infrared image heating furnace, where the temperature was rapidly raised to 1370°C at a rate of 1000°C/min in a nitrogen stream, and then held at that temperature for a predetermined time before rapid cooling. Following rapid cooling, the internal structure of the sample was examined using a high-resolution X-ray CT device. The spherical equivalent volume was calculated based on the number of observed bubbles and their equivalent circle diameter. Additionally, the relationship between the volume ratio of small bubbles in the slag volume and the distance from the bottom of the crucible was determined, along with the calculation of bubble density and volume ratio. Under various experimental conditions, numerous bubbles with equivalent circular diameters ranging from 200 to 300 μm were consistently observed. The density and volume fraction of bubbles with circular equivalent diameters between 200 and 500 μm showed a tendency to increase as the distance from the bottom of the crucible increased. Bubbles with equivalent circular diameters of 500 μm or more consistently exhibited lower bubble densities, regardless of their height within the crucible. The formation of larger bubbles is attributed to a decrease in iron oxide concentration within the slag as the reaction progresses. However, this is believed to be due to their higher buoyancy, resulting in shorter residence times within the slag and a reduction in the number of encapsulated bubbles.

Aug 21, 2024

By M S. Kizil, T OÆConnor

It is predicted that underground coal mines in the Bowen Basin will have a significant rise in heat problems as mines aim to increase production, depth of extraction and length of longwall panels. At an actual Bowen Basin mine, extended longwall panels greater than 4 km in length are planned to be mined at depths with a virgin rock temperature (VRT) of more than 40¦C. Research has been completed into the reasons for the substantial increase in temperature at the face for underground coal mines in the Bowen Basin. An analysis into heat loads produced at the Bowen Basin mine for the development and production phases shows a significant amount of additional cooling will be required to ensure safe working temperatures. All of the major sources of heat, which included both natural and mine induced heating, were considered for the analysis. Maximum heat loads were assessed, together with peak intake surface temperature to determine the maximum cooling capacity required at a Bowen Basin mine. These maximum heat loads were assessed for a specific longwall panel, Longwall A, which extends to over 4.1 km in length and at a maximum depth of over 330 m, resulting in a VRT above 40¦C. The cooling system positional efficiency was analysed, determining that underground cooling via an indirect heat exchanger was preferred to increasing the capacity of the current bulk air cooling (BAC) on the surface. A preliminary cost analysis was also detailed which compared the cost of additional BAC compared to implementing underground cooling.CITATION:OÆConnor, T, Kizil, M S and Belle, B, 2013. Required cooling system capacities for extended longwall panels at a Bowen Basin mine, in Proceedings The Australian Mine Ventilation Conference , pp 249-256 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 1, 2013

By M Nagarajan, S Esen

This paper presents three case studies from three surface coalmines. The first case study is from a coalmine where a dragline operates in 60 m wide and 1 km long strips. Typical overburden height ranges from 30 m to 50 m. The base case blast was modelled using an advanced blasting model (Distinct Motion Code – DMC) to establish the baseline using the blasting parameters in use before optimisation. DMC was used to model the alternative blasting scenarios to obtain the following key objectives:the post-blast muck pile profile was required to have an optimum height of ~28 moptimise cast percentageachieve a well-fragmented muck pile for the draglinethe muck pile profile should be appropriate for the dragline entry into the strip.Achieving these objectives would optimise the dragline performance through reducing rehandle and improving advance along the strip. DMC modelling resulted in the following outcomes:the required muck pile profile could be achievedthe cast percentage increased from 21.1 to 25.1 – a four per cent increaserehandle volumes were reduced.Establishing a good quality assurance / quality control (QA/QC) process was essential to delivering a blast that could achieve these outcomes. The blast was fired successfully and the dragline completed the strip approximately two weeks ahead of the time taken for the previous strip.The second case study is from a coalmine where D11 dozers are used to move the blasted material to the spoil. An optimum profile was required for the dozers for this low-cost mining operation. To achieve this, an electronic blasting system (uni tronicTM 600) was introduced to optimise the muck pile profiles being delivered. The new muck pile profiles showed improved cast percentage from 10.6 per cent to 17.8 per cent. The changes in the muck pile centre of mass displacement improved dozer efficiency as a result of pushing material shorter distances to its final position. Dozer productivity (bcm/h) statistically did not change but coal was being exposed at a faster rate.The third case study summarises the results from a site where dozers are used to minimise the total mining cost. Traditionally, excavators had been used on-site with conventional blast designs. After firing the first cast shot, the blasted material was found reasonably well-fragmented. Although dozing productivity hasn’t been quantified on-site, operators were providing positive feedback (easier push) due to the reasonably good muck pile profile suiting to the dozers. This site also realised significant savings with the dozing as compared to their truck/shovel operation.CITATION:Esen, S and Nagarajan, M, 2015. Muck pile shaping for draglines and dozers at surface coalmines, in Proceedings 11th International Symposium on Rock Fragmentation by Blasting, pp 409–416 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 24, 2015

By H Saw, M Cordova

Cemented rock fill (CRF) is often used to support open stope voids. This allows for the full recovery of ore while achieving global mine stability. In such cases, the exposed CRF masses require adequate compressive strength and stiffness to resist the forces and limit the displacement associated with movements in the rock mass surrounding the excavations. The CRF material preparation often involves the use of large particles, either from crushed rock or development mining waste. This means that conventional laboratory samples cannot be readily used to determine the laboratory strength. Consequently, a research project was undertaken to determine the influence of sample size on uniaxial compressive strength. This paper presents the results from CRF samples that were prepared using the same mix design and then cast into moulds with diameters of 150, 240, 300 and 400 mm. The laboratory testing also allowed a better understanding of the effect of particle size distribution upon the overall strength of a rock fill mass. The laboratory results were compared to a database of large-scale results from the testing of fill masses from a number of mine sites.CITATION:Cordova, M, Saw, H and Villaescusa, E, 2016. Laboratory testing of cemented rock fill for open stope support, in Proceedings Seventh International Conference and Exhibition on Mass Mining (MassMin 2016), pp 755–762 (The Australasian Institute of Mining and Metallurgy: Melbourne).

May 9, 2016

By N Cusworth, W R. Mackenzie

"In 2007, the authors presented a framework for the conduct of resource project feasibility studies and provided guidance on minimum standards and best practice (Mackenzie and Cusworth, 2007).This paper will present the authors’ observations in relation to feasibility studies undertaken to support investments made during the unprecedented period of resource project capital investment since 2007. It will also provide comments on general trends in the mining industry’s approach to project assessment and update recommendations on best practice.CITATION:Mackenzie, W R and Cusworth, N, 2016. The use and abuse of feasibility studies – has anything changed?, in Proceedings Project Evaluation 2016, pp 133–148 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Mar 8, 2016

By Yates C, Gillie E. W, Finucane K. J

During the year 1940 a complete investigation was made of the whole of the stoping operations throughout the Great Boulder Mine and, as a result, it was decided to convert some of the existing shrinkage stopes into cut-and-fill stopes, and also to mine many of the more recently developed ore blocks by the latter method. The necessity for filling was due primarily to the structure of the ore blocks, the character of the walls and possible excessive dilution, but was influenced also by the grade of the ore-block and by the fact that cleaner mining can be carried out by cut-and-fill stoping.The comparatively few cut-and-fill stopes in operation prior to 1940 had been filled mainly with waste rock from development work, but as the amount of this material available, or likely to become available, would have been wholly inadequate for future operations, it was proposed to construct a system of passes to supply filling material from the surface.Owing to the distribution of the stopes, no central pass or system of passes could be devised which would service all the stopes satisfactorily.However, work was commenced on two passes, one designed to service the cut-and-fill stopes in the Eastern Section of the mine, and another to service those in the Western Section. It was realised that these passes could serve only a small proportion of the total number of Stopes (Figs. 1 and 2). During the progress of this work experiments were undertaken to test the possibility of using filtered tailings underground as an adjunct to the above scheme, but for various reasons these were only partially completed.Early in 1942 it was suggested by the Mill Superintendent that an attempt be made to fill with wet classified flotation tailings as practised at Mt. Lyell in Tasmania and at the Triton Mine at Reedy, West Australia. For the purpose of testing this, an experimental stope was prepared on the 309-ft. level, "X" Lode. This was 100 feet long and averaged 4 feet in width. Water-tight rearings were built at each end of the stope and waste rock fill was spread to a height of 6 to 8 feet above the timber. Used filter cloths were laid on top of the fill and brought a few feet, up the walls of the stope. A two-inch victaulic drainage line was then placed in position, the end passing through a hole in one of the bearings and thence to a sump. In the central portion of the stope two vertical pipes were brought up from this line. These were bored with one-inch holes staggered on opposite sides of the pipe, the vertical interval between each hole being one inch. The holes were then stopped with...

Jan 1, 1945

By J Fisher, M Castro, A Hariyadi

The train haulage system being designed for PT Freeport Indonesia’s Grasberg Block Cave (GBC) will be located within the mine at an elevation of almost 3000 m above sea level on the west side of the island of Papua, Indonesia. The developing block cave is beneath the active Grasberg Open Pit and will transport 160 000 t of ore per day from multiple chute galleries to the underground crushing plants. The design entails 28 km of rail and 119 ore chutes, which will be brought online over the course of ten years.The train haulage system will be driverless and fully automated. Sequentially, the system includes remote controlled rock breakers at the top of the chute grizzlies, followed by remote controlled chute loading of the trains, then autonomous train haulage and autonomous unloading. An overhead catenary system (OCS) will power the locomotives, which are equipped with on-board batteries to drive through the loading and unloading sections. Service locomotives with drivers move material-laden freight cars in and out of the service shaft, destined for workshops, the magazine and batch plants. Service locomotives are also used for haulage level maintenance. Highly reliable and redundant control and communications systems are required to manage 12 trains in simultaneous operation.This paper provides an overview of the train haulage system’s technical complexities and the associated support systems interactions. The paper also highlights the project management issues and successes of this globally sourced project.CITATION:Hariyadi, A, Castro, M and Fisher, J, 2016. Automated train ore transport, in Proceedings Seventh International Conference and Exhibition on Mass Mining (MassMin 2016), pp 563–570 (The Australasian Institute of Mining and Metallurgy: Melbourne).

May 9, 2016

By D R Quick

The Goodall gold mine is located 30 km east of Adelaide River, within the Pine Creek 1:250 000 map sheet, at 8 525 000 mN: 750 000 mE, on the AMG grid. The first gold was poured in April 1988 and the mine closed in September 1993, having produced 228 400 ounces of gold from 4.095 million tons of ore, at a head grade of 1.99 g/t. Oxide and primary ore was mined from one large open cut and low grade oxide ore from three smaller pits. The gold mineralisation was discovered in 1981 by means of a helicopter borne rock chip sampling survey, which delineated 13 discrete anomalies. Systematic grid sampling of rock chips outlined an area of 1300 m by 300 m within the 2.0 g/t contour. Over 7500 m of costeaning was used to delineate the A and B-Pod mineralised areas. Geochemical soil sampling played only a minor role in the discovery and geophysics none at all. By the end of 1984, 40 diamond drill holes had been completed and a resource of 3.63 million tonnes at 3.0 g/t calculated. In January 1985, Western Mining Corporation Ltd became a joint venture partner and funded exploration to cam a 60 per cent interest. Additional drilling led to a feasibility study in 1987, based on a Measured and Indicated resource of 4.25 million tonnes at a grade of 2.35 g/t. The host rocks are quartzo-feldspathic wacke and interbedded phyllitic siltstone, of the Proterozoic Burrell Creek formation. The main zone of gold mineralisation, A-Pod, occurs in a sub-vertical vein array, up to 50 m wide and 800 m in length, of quartz and sulphide veins (5-20 mm wide), It is parallel to and on the east limb of an upright, north plunging anticline. Five types of veining have been recognised, with the bulk of the gold mineralisation associated with quartz-pyrite-arsenopyrite and chlorite-quartz-pyrite veins. The gold occurs as particles, from 1-20 micron, as inclusions and along fractures in pyrite and arsenopyrite and to a limited extent in chlorite. The mineralisation is epigenetic and structurally controlled by fractures developed during an earlier folding event. The gold is assumed to have been derived from the underlying sediments of the Koolpin Formation by hydrothermal fluids emanating from the deep seated northern margin of the Cullen Batholith. Precipitation of gold may have been controlled by variations in temperature and pressure.

Jan 1, 1994