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By G Bickert

Australia is currently developing and engineering a range of specialised processing options for rare earth (RE) (Lynas, 2011, Alkane, 2005), zirconia (Alkane, 2005), lithium (Galaxy Resources, 2008) and other special minerals and metals required in high-tech applications. The processing of lithium, zirconium and RE minerals is based on complex leaching processes which require extensive liquor clarification, solid liquid separation and solids washing. The filtration properties for those leach liquors, residues and precipitates are not well understood and differ drastically according to the process route applied. Thus, filtration test work is essential for equipment selection, sizing and design. Similar types of slurries are filtered in laterite nickel, TiO2 and other chemical oriented processes, while more experience is available here as those processes are applied widely for a number of decades. The increasing need for dry stacking of tailings requires advanced tailings filtration with properties differing drastically depending on tailings properties. More traditional mineral processes such as flotation commonly apply filtration. All of these and other cases require proper filtration test work particularly evaluation of vacuum versus pressure filtration and the need for cake pressing and/or air blowing for filter presses. Bench scale test work and/or pilot plant operation is common to determine filtration rates, achievable cake moisture and wash efficiencies and sometimes filtrate clarity. While those values are important, there is much more valid information to be obtained by filtration test work.This paper discusses the intricacy of filter selection and sizing based on driving forces (vacuum and pressure air, compression by mechanical means and slurry) and parameters (cake thickness, forces and times) and different processes utilised in common filtration equipment. It will discuss the in’s and out’s of filtration test work design, planning, and in particular data gathering, processing, reporting, evaluation and filter sizing from recent examples of bench and pilot scale filtration test work for above applications together with methods for predicting filtration performance for example where test work is lacking. One focus will be the realistic extrapolation of test work results towards different parameters utilised in full scale equipment and plants. The conclusion of the paper will contain checklist for filtration test work.

Aug 8, 2011

By D Connelly

Quartz deposits are common in nature but the presence of quartz suitable for yielding high purity quartz is exceedingly rare. This paper describes a test work study on producing high purity quartz (HPQ) product to meet the tight product specifications of greater than 99.97per cent, including four and even five nines, for uses including, lighting and fibre optics and modern-day high efficiency solar panels. Due to demand in renewable energy applications the market for high purity silica solar panels known for their increasing efficiency and durability is rapidly growing. Demand continues to rise for the use of high purity quartz in semiconductors, electronics, imaging and sensor markets. Microelectronic components are created by chemically fabricating wafers of semiconductors such as silicon to obtain the desired transport of electrical charge and control of the current. The production of silicon metal from high purity quartz has been designated a Strategic Mineral by the European Commission and by the US Department of Justice due to the growing significance of this metal in various industries. The subsequent processing of quartz is the key step to adding value and aiding technological advances. Overall, the high purity quartz market is a highly secretive, specialised market with increasing opportunities for stakeholders. The attempt to produce HPQ in this work was carried out using an Australian quartz ore following a typical HPQ process flow sheet. The unit processes included primary crushing, scrubbing, magnetic separation, flotation, acid leaching followed by hot chlorination. The results indicate that grinding of quartz requires specialist equipment due to the hardness and extreme abrasiveness of the ore leading to additional iron contamination in the ground ore. The critical impurities were aluminium, iron, sodium, potassium, lithium, titanium, zircon, calcium and magnesium. The chemical refining was found effective. The hot chlorination process needed to be done at 1200°C in a chlorine – hydrogen chloride gas atmosphere. The risk of any gas being released during this stage is very serious and only specialist companies should and can carry out this step. More recent studies reported many new options that appear to facilitate the production of HPQ with the requisite purities more easily, but costly equipment along with highly trained operators are required and hence unlikely to become available in standard metallurgical laboratories any time soon.

Sep 1, 2024

By C Lower

An unprecedented, quantitative, deposit-scale understanding of major clay mineral distribution throughout the Yeelirrie uranium deposit has been achieved by adapting the existing HyLogger™/HyChips™ analytical technique to measure resource assay pulps. Spectral analysis via HyLogger™ and HyChips™ systems has previously been used as a small-scale geometallurgical research tool, but never before has it been quantitatively applied to an entire resource sample population for the purpose of developing stand alone block models for major clay minerals. Economic uranium mineralisation within the thin but extensive Yeelirrie?palaeochannel?primarily exists within the transition zone between a groundwater-derived calcrete and the underlying clay-quartz regolith. Early geometallurgical investigations identified that certain clay minerals could be deleterious to the efficiency of the uranium recovery process. Accurate calculation of clay mineralogy from resource assays proved too difficult due to chemical similarities between co-existing kaolinite, smectite and magnesium carbonate minerals. A relatively low-cost, efficient and accurate method of quantitative clay mineral analysis was required to measure over 70 000 resource assay pulps. After initial testing and comparison with QXRD, the HyChips™ 3-1 spectral logging system was implemented as a primary method of clay mineral analysis for the Yeelirrie uranium deposit. The resultant spatial data set allowed the interpretation and geostatistical analysis of geometallurgical domains across the deposit. Clay block models were generated and used in geometallurgical resource calculations to benefit the mine planning process.

Sep 5, 2011

By J deVries, A Broodryk, P Kyselica, K McLean

Gwalia gold mine is located close to the town of Leonora in the Goldfields region of Western Australia, some 830 km from Perth. The mine is wholly owned by St Barbara Ltd, an Australian Securities Exchange listed gold producer and explorer. The Gwalia mine is a deep, all-trucking mine and is serviced by a 10 km long decline from surface. Mining is currently 1540 m below surface. Reserves extend to 1820 m below surface and are open at depth. The mine produces in excess of 200 koz of gold per annum.The Leonora area is hot and arid, with a mean annual rainfall of 230 mm and a 95 per cent mean daily maximum temperature of 20.3°C wet-bulb and 35.6°C dry-bulb.The primary ventilation system comprises a 490 kg/s surface exhaust fan, which draws air down a parallel intake system consisting of the decline, and a dedicated fresh-air intake system extending to the bottom of the production zone. Cooling is provided via a surface bulk air cooler serviced by an ammonia refrigeration plant that delivers 150 kg/s of chilled water to the bulk air cooler.The refrigeration plant and bulk air cooler were installed in 2008. The design allows cooling to be increased over time by installing an additional refrigeration plant in a lead lag configuration with the current refrigeration plant when required.Since 2008, expansion and increased depth reached a point where the underground heat load exceeded the installed capacity of the existing 4.0 megawatt rated (MWR) refrigeration plant. In order to maintain the mine design conditions underground, upgrading of the existing refrigeration system was required to bring the cooling capacity to 7.5 MWR.Two expansion options were considered: the addition of an ammonia vapour-compression refrigeration machine similar to the existing one or a lithium bromide absorption refrigeration system using waste heat from the site power station. The original design and layout of equipment on-site meant that either could be used, allowing an assessment of the merits of both when additional cooling was required.The decision was made to install a lithium bromide absorption system on the lead side of the existing ammonia refrigeration plant. This arrangement provides considerable economic benefits, including reductions in site-generated electrical power, generator fuel consumption and carbon dioxide emissions.This paper details the planning, design, installation and commissioning of the absorption chillers.CITATION:Broodryk, A, deVries, J, Kyselica, P and McLean, K, 2015. The design, installation and commissioning of an absorption refrigeration system at Gwalia gold mine in Western Australia, in Proceedings The Australian Mine Ventilation Conference, pp 193–200 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 31, 2015

Standard petrographic profiles for many Permian seams from New South Wales and Queensland have been drawn up. The standard profiles of seams of the same age and from one or more coalfields were used to compile an idealized or type seam sequence for those coals in which more than half the seam profiles were similar to one another. A marked similarity existed between at least 7 of the 10 type seam sequences formed. The configurations of the vitrite-plus-clarite contents of the plies of 95 per cent of the seams could be formed from different arrangements of 3 basic sequences: (i) a two-ply sequence with a lower vitrite-rich and an upper vitrite-poor ply; (ii) a two-ply sequence with a lower vitrite-poor and an upper vitrite-rich ply; and (iii) a four-ply sequence of 3 plies with vitrite-plus- clarite decreasing upwards, terminating in a vitrite- rich ply. The third basic sequence was very similar to a succession described for some British Carboniferous coals. Eighty per cent of the Australian Permian coal seams studied began with the first (i) basic type of sequence and are interpreted as originating in peat- forming areas which became progressively drier. More than half the seams were formed in an area with a stable depositional surface level which was subject to oscillations of small amplitude.. Many (30%) were formed in an area with a steadily rising depositional surface level.

Jan 1, 1969

By P Bartsch

Uranium solvent extraction, USX has been applied commercially for recovery and concentration for over 50 years. Uranium in acidic liquor, which is prepared following ore leaching, solid/liquid separation and clarification, can be treated through a sequence of operations; extraction-scrubbing-stripping, to obtain purified liquor. USX has dominated the primary uranium industry as the preferred technological route to converter grade yellowcake. The practices of design and operation of USX facilities has found renewed interest as new mines are developed following decades of industry dormancy. This article seeks to outline principles of design and operation from the practitioner’s perspective. The discussion also reviews historical developments of USX applications and highlights recent innovations. This review is hoped to provide guidance for technical personnel who wish to learn more about good practices that leads to reliable USX performance.

Aug 8, 2011

By Smallnone P, Walker B

Mining of the AKI lamproite pipe at the Argyle Diamond Mine (ADM) commenced in 1985. Originally designed to mine ore at a rate of 3.6 Mt per annum, the project has gone through two major plant expansions to a designed production rate of 8.7 Mt

Jan 1, 1998

In the current climate, there is a real need to bridge the gap between theory and the real world mining considerations, and develop viable, cost-effective, high performance thickeners. With a greater need to extract profits from both difficult-to-process ores and lower grade orebodies, coupled with the necessity to produce 'greener', less water intensive technologies, suppliers need to take current thickener designs to the next level and into the 21st century. The complexity of 'getting it right' within a thickenerÆs feedwell, the 'nerve centre' of the thickener, is no small task. Delivering dilution water and mixing this water with incoming feed slurry, addition of flocculant, maximising flocculant-slurry interaction to build aggregate size, minimising unabsorbed flocculant percentage, volume average of solids volume fraction, momentum dissipation ratio, dilution recycle ratio, volume average shear, maximum shear rate at exit, aggregate size on exit, uniform discharge profile and more ù all of these need to occur in the right order and in the right proportions within the feedwell. This paper discusses the genesis and evolution of a new feedwell design, the first major innovation in the thickening industry in over 15 years.

Jan 1, 2008

Flat jack stress determinations in selected colliery roadways have revealed a range of measured stress from 330-p.s.i. tension to 3,750 p.s.i. compression. Photoelastic examination of loaded plastic and actual rock mo.odel mine openings has shown that the primary horizontal stress required to attain these measured stresses would be from one to three times the superincumbent vertical load.Examination of roof movement with a. borehole oxtensumeter and a series of anchors at various horizons to 30 feet in a. vertical roof borehole showed that when a standing face waa.s advanced the roofmovement was practically confined to the lower two feet. The ma.j ority of this movement occurred during the 36 hours after mining at this face commenced.The use of a. radioactive isotope and a. scintillation counter in a. pair of parallel bcreholes promises to be sensitive and accurate method of roof bedding plane crack detection.

Jan 1, 1963

By X Sun, L B. Zhou, H Yi, X Tang, P Ning

The rapid industrialisation in China has consumed large amounts of various industrial raw materials, which has left behind large quantities of industrial solid wastes produced from the processes of mining, mineral processing and smelting. The storage of huge quantities of industrial solid wastes is an obvious risk to the environment and human well-being. As the first step to fully appreciate the scale of this risk, it is necessary to carry out research for developing a pollution source environmental risk assessment system (PSERAS), especially for the typical staple industrial solid waste (TSISW). This paper aims to discuss the frame of the PSERAS for TSISW for large scale production. This system consists of two aspects: predicted factors and accidental ones, which involve physical, chemical and biological risk assessments, on-site and during transportation process. Two analysis methods, the analytic hierarchy process (AHP) and fuzzy logic assessment (FLA) were used to handle the complexity of environment and non-quantitative data. The risk assessment system for such a complex objective is difficult and the assessment of relative factors is problematic through the usual quantitative analysis. With the proposed PSERAS, it is possible to define the risk class of pollution sources. Furthermore, the risk assessment results will be helpful to the decision-maker, especially when facing unknown risks in the environmental management of industrial solid wastes. In addition, by means of the PSERAS, risk factors can be ranked against their contributions to environmental risk, in order to calculate their relative priorities in decision-making.CITATION:Sun, X, Ning, P, Tang, X, Yi, H and Zhou, L B, 2012. Environmental risk assessment system for typical staple industrial solid wastes, in Proceedings Life-of-Mine 2012 , pp 301-304 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 10, 2012

By Wasserman A. D

Methodological approaches and in- vestigation results are considered in the paper of heatmass transfer processes when harmful impurities are scattered and remo- ved from chamber-like workings, of the air distribution in complicated ventilation sys- tems and of chemical conversion of exha- ust gases in mine atmosphere when diesel equipment is operating. The results presented in the paper are the base for creation of the automatic design, control and regulation of the venti- lation processes.

Jan 1, 1988

The Comstock Lode, a world-class precious metal vein district near Virginia City in western Nevada, U.S.A. (Fig. 1), produced 8.3 million ounces (0.28 X 106 kg) of gold and 193 million ounces (6.56 X 106 kg) of silver during the period 1859 to present (Bonham and Papke, 1969). More than 95% of production was from the Comstock fault, while the adjacent Silver City, Occidental and Flowery vein systems produced about $10 million (U.S.). Within the Comstock fault high-grade ore bodies, or bonanzas, were distributed over 2.3 miles (3.8 km) along strike and persisted to >3000 feet (>900 m) below the surface (Fig. 2). Bonanzas consisted of as much as 1.4 million tons (1.3 million tonnes) of ore grading up to 1.9 oz./ton (65 gm/ton) Au and 30 oz./ton (1020 gm/ton) Ag (Table 1). Some bonanzas occurred in the main Comstock fault plane while others, including the highest grade bonanzas, formed in hangingwall frac- tures. After 1880 most ore consisted of lower-grade vein material surrounding mined- out bonanzas and large volumes of mineralized vein and wall rock in the Comstock hanging- wall that were mined from open cuts. Deep mining was hindered by large quantities of hot water and bad ground which ultimately defined the limits of underground production. The widths and richness of Comstock bonanzas strongly influenced U.S. mining law, milling and underground mining methods, and securi- ties regulations that are in practice today. Within the district Mesozoic metasedimentary rocks, metavolcanic rocks, and granitic in- trusions are overlain by Miocene volcanic rocks (Thompson and White, 1964). Veins cut all lithologies. Miocene volcanic rocks and their K-Ar ages (Vikre, McKee and Silberman, 1987) include Santiago Canyon rhyolite (22

Jan 1, 1987

By N Jamson, T K. Rohde

The Mine is an underground zinc (Zn), lead (Pb) and silver (Ag) operation in central New South Wales (NSW). Broadly speaking the Mine has two domains that require rehabilitation, the mine industrial area (MIA) and the tailings storage facility (TSF). The MIA has been used for the handling and processing of ore and contains elevated sulfate, sulfide and metals (Pb, Zn). The TSF receives the by-products of processing, that is, potentially acid forming (PAF) tailings which also contain elevate sulfate, sulfide and metals. Typically, the rehabilitation of a TSF involves two controls. Firstly, controlling the potential for PAF tailings to form acid rock drainage (ARD) by limiting interaction with oxygen. Secondly, once the tailings are unsaturated, limiting interaction with water (usually as rainfall) to reduce the potential for ARD to be transported to the receiving environment. These two controls are often employed in the form of a cover.In semi-arid environments there are many Australian examples of covers that have been built to limit interactions of mine by-products with oxygen and water. Typically, the covers contain one or more layers and are built from soil and rock which may include run-of-mine rock. The Mine however has limited supplies of stockpiled soil; and run-of-mine rock is not generated in large volumes because it is used as backfill underground. A cover design strategy was developed to cater for these limiting factors. The strategy identified soil, ore and rock accretion (the accretion) below the MIA could be a potential source of borrow material for the cover. The purpose of this paper is to describe the cover design process, the method and results of large (2 m tall) column trials that were used as a cost-effective way to trial three cover options and finally, the paper describes the water balance that was developed from the column trial results and subsequent modelling using SVFlux.CITATION:Rohde, T K and Jamson, N, 2018. Making use of potentially contaminating mine waste in cover design, in Proceedings Mine Waste and Tailings Stewardship Conference 2018, pp 285–298 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 23, 2018

By P Holtham, K Duffy, L Pyle, W Valery

The mining industry is becoming increasingly focused on pre-concentration due to its potential economic and environmental benefits, particularly for the more complex, low-grade, high throughput operations now being developed. Advances in technology have seen bulk ore sorting (BOS) trialled at several operations. However, BOS should not be seen as a ‘silver bullet’, it can offer significant benefits, but only under the right conditions. Deposits with interspersed mineralogy or disseminated, fine-grained minerals can be difficult for BOS. Large batches of barren or low-grade material are required for BOS to be effective. Hatch has recently completed several studies for clients who believed that bulk sorting was the most suitable pre-concentration technique for their operation, based only on an assumed sorter recovery or on laboratory trials of the chosen sensor. However, minimal consideration had been given to the heterogeneity of the deposit, the achievable separation batch size or sensor repeatability. Evaluation of BOS needs to consider both heterogeneity and sensor suitability. The valuable or gangue mineral must be present in such a way that it can be separated effectively in the large bulk sorting batch sizes, without competing minerals or changing ore characteristics affecting the metal upgrade or recovery. Also, the sensor chosen must be able to detect either the valuable mineral, a proxy for the valuable mineral, or a gangue mineral for rejection with sufficient accuracy and precision (repeatability). This paper presents three recent case studies, emphasizing markers which indicated that bulk sorting was unsuitable for the project whilst highlighting the pre-concentration technologies which were applicable (including particle sorting, and coarse particle gravity separation). The modelling methodology, equipment sizing and forward work plans developed to verify these alternative solutions are also presented, demonstrating the key factors contributing to a thorough scoping assessment.

Aug 24, 2022

By Warner RE

SINCE 1952, when Elliott et a/ first applied the Sherwood correlation for flooding in chemical absorption towers to blast furnaces numerous investigators throughout the world have followed their example and much research effort has been spent to tighten the correlation and recently to extend investigation of flooding to side gas entry conditions. The widespread use of the Sherwood concept of flooding has been such that even the word itself has become an integral part of blast furnace nomenclature. The evidence for the applicability of Sherwood's relationship has been that the flooding limits calculated from blast furnace parameters lie approximately on the Sherwood line. However, the values of many of the variables for the blast furnace calculation are only estimates and at first sight one may wonder at such 'close' coincidence when some of the estimated parameters entering the Sherwood correlation are raised to the power of three or more, as for the bed porosity and the liquid surface tension. It is felt that the explanation of this result lies in the logarithmic: nature of the plot and/or the fortuitous can- cellation of the errors of the estimates in the calculations coupled with the fact that blast furnace fluid ratios lie within a very narrow range of values. The foregoing considerations should in their own right place doubt on the correctness of the Sherwood type flooding correlation applied to the blast furnace. Additional doubt regarding the existence of the Sherwood type flooding in the furnace has been cast by the results obtained in the Tokyo University experimental blast furnace. From the reasoning of both Standish and Colquhoun (p. 20) and Warner (p. 23) it is deduced that flooding in the traditional sense will only normally occur in the upper bosh (somewhere near the mantle), following softening as discussed by Palella et a/ on p. 75, and flooding in that zone is likely to result in furnace irregularity. On the other hand, the concept developed by Warner of slag hold-up above the raceways as a stable phenomenon not affecting furnace regularity is a process unique to the blast furnace and is not described by the Sherwood flooding theory. Recognising the considerable practical implications of slag hold-up above the raceways, this note further explores the physical process of slag hold-up. We envisage that bed of coke above the raceways in almost all blast furnaces would contain some slag held up by the gas pressure gradient. In general the slag would be contained in a ring and gas would bubble through in its ascent through the furnace. The force supporting this slag is the pressure gradient and obviously -much of the blast pressure is dissipated in supporting this structure.

Jan 1, 1975

By M Saghaeian, A H. Baladastian, A Dehghani Ahmadabadi, V Sheikhzadeh, R Hejazi

This research was done to reduce silica from mixed hematite concentrate of Barfkar Iron Ore Sample and increase Fe grade using reverse flotation method (Depressing Hematite and floating Silica). The used sample in the reverse flotation tests was a mixture containing two types of hematite concentrates (Spiral cleaner concentrate and Slon-VPHGMS cleaner concentrate) with the percentage of 16 per cent and 84 per cent, respectively. Fe grade of two samples was 59.31 per cent and 57.52 per cent and SiO2 grade was 5.83 per cent and 6.71 per cent, respectively. After mixing, the Fe and SiO2 grade of the mixed sample was measured at 58.10 per cent and 6.20 per cent. For flotation tests, effect of variable operational parameters including feed size distribution (d80 = 130 microns and d80 = 53 microns), pulp solidity (25 per cent and 30 per cent w/w), pH (9.70, 10.20 and 11), Sodium silicate as a dispersant (500, 1000 and 2000 g/ton), Starch as a depressant (500, 1000, 2000 and 3000 g/ton) and EDA as a collector (180, 300, 400 and 500 g/ton) were investigated. For evaluating the influence of Sodium silicate and Starch dosage, initial flotation tests were performed on the main sample (d80 = 130 microns). These two parameters had no significant effects on the yield, Fe, and SiO2 grade of flotation concentrate based on test results. In the next step, screen analysis was carried out on the main sample for mineralogical studies. Accordingly, the suitable size distribution for releasing Fe and SiO2 particles was smaller than 53 microns. Therefore, the sample was crushed until d80 = 53 microns. Then, supplementary tests were performed on the crushed sample for studying the pH range, pulp solidity, and EDA collector dosage. After doing silica removal flotation tests based on the designed pattern and comparison of the results, the conditions consisting of size distribution d80 = 53 microns, pulp solidity 25 per cent, pH 10.20, Sodium silicate 1000 g/ton, Starch 1000 g/ton, and EDA 500 g/ton was selected as optimum values. In the last test, Fe and SiO2 grade was measured at 60.00 per cent and 3.93 per cent, respectively, and Fe recovery was obtained at 74.93 per cent.

Aug 24, 2022

The conventional approach to the disposal of fine coal mine tailings, which involves piping the tailings as an aqueous slurry to a storage where they remain in the form of a permanent wet deposit, is becoming unacceptable. This approach is a legacy of a "contain and forget" mentality, which has economic advantages only in the short term. In the longer term, wet tailings deposits occupy large areas of land and present extreme rehabilitation difficulties. Alternative tailings disposal systems, which are sustainable and which facilitate rehabilitation, are required. The possible alternatives include the mechanical dewatering of fine tailings prior to disposal in combination with coarse reject, tailings disposal into spoil pile valleys, and the combined pumping to disposal of tailings and coarse reject. The most appropriate choice of fine tailings disposal strategy will vary from mine site to mine site. The chosen strategy should be optimised to meet increasingly stringent environmental requirements, within the constraints imposed by each mine. The alternative approaches represent significant departures from past practices, and require considerable industry driven research and development before they can be adopted with confidence. All three alternatives involve mixing of the fine tailings and coarse waste streams, aimed at creating a mixture of reasonable engineering properties. Understanding the engineering behaviour of such mixtures is central to the success of these alternative disposal strategies.

Jan 1, 1991

By F Musa

One of the main energy efficiency opportunities identified from the majority of mine site energy audits is to optimize the comminution circuit to reduce energy consumption through improved throughput and product size distribution. Typical analysis reviews comminution process efficiency/inefficiency i.e. utilization of grinding mills, process bottlenecks and product vs. target size, to identify instances of overgrinding or undergrinding. From this, it is possible to establish the potential improvement that could be gained through formal plant optimization and what the optimization program would entail. It can also sometimes reveal immediate opportunities for process improvement. This paper presents steps that can be taken to optimize the efficiency of comminution circuits.

Aug 8, 2011

By D J. Kilkenny, D A. O’Toole, P B. Hills, K L. Seah

An investigation was undertaken to critically analyse the use of fibre reinforced shotcrete (FRS) for a variety of applications in large diameter shafts. Those applications included the use of FRS with or without the addition of subsequent rock bolting in the temporary support cycle prior to the installation of a permanent lining, most likely cast in situ concrete or as the final lining medium.Numerous authors have considered the early re-entry and strength characteristics of FRS in recent years. The current study has investigated the practical application of the recommendations of many leading authors in an endeavour to establish practical and acceptable risk-assessed procedures for the use of FRS for rapid, safe, efficient and cost-effective shaft sinking.Shaft sinking is regarded as a specialist and potentially higher risk activity due to the confined work environment, exposure to ground conditions and a high degree of manual handling associated with getting materials to the sinking face. The use of robotically applied FRS as a primary lining in shaft sinking has significant safety advantages in reducing manual handing associated with mesh safety support and bolting, as well as reducing direct exposure to ground conditions.The study involved both field testing and a literature review. The study established an initial set of criteria for the use of FRS in the development of large diameter shafts and demonstrated a number of scenarios where the technique of robotically applied FRS provided significant safety and cycle time advantages. The opportunity for a fully sprayed single pass lining was also investigated as part of the project.CITATION:Hills, P B, O’Toole, D A, Kilkenny, D J and Seah, K L, 2014. Exploring the use of fibre reinforced shotcrete in large diameter shafts, in Proceedings AusRock 2014: Third Australasian Ground Control in Mining Conference , pp 295–306 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 5, 2014

By P Edmondson, S Narendranathan, S Setchell

The complexities of mining in current economic conditions, where ore extraction schedules are varied on a weekly basis so as to ensure constant supply to the process plant, can result in a number of scenarios that pose geotechnical challenges. This paper presents a case study where alterations in the mining schedule of a particular mine resulted in the necessity to mine under a large uncompacted embankment, ie integrated tailings and mine waste landform. To effectively ensure that ore recovery was maximised, a well-designed slope monitoring system to evaluate design performance and to provide timely warning of impending failures was imperative. In order for an appropriate slope monitoring strategy to be developed, a thorough understanding of the potential slope failure mechanisms, capabilities and limitations of available monitoring techniques is required. Communication of the monitoring results and response strategy to alarms generated by monitoring systems must also be carefully considered to maximise effectiveness. The results of preliminary stability modelling can be used to guide the design of a monitoring system, which in turn can then generate data which can be used to iteratively refine the geotechnical model. This paper illustrates the development and implementation of a slope monitoring system and associated communication system, using a case study developed at a large iron ore open pit mine in the Pilbara region of Western Australia.CITATION:Narendranathan, S, Lee, E C, Setchell, S and Edmondson, P, 2015. Modelling of an integrated waste storage facility, in Proceedings Tailings and Mine Waste Management for the 21st Century , pp 183–190 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 27, 2015