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Zircon is extensively used as moulding sand in the production of steel and high-duty iron castings. To achieve a satisfactory grade of zircon concentrate, the zircon particles may need ,to have surface coatings removed, as such coatings can hinder electrostatic and magnetic separation of zircon from associated minerals. Xn addition the presence of surface coatings on the zircon. particles can cause sinterins and fusion of the zircon to the east metal resulting in "burnt-on" sand or fritting of the casting. "Burnon" necessitates cleaning of the casting by grinding or chipping, which is expensive and results in loss of definition on the casting face.Hudson (1966) has pointed out ,that some coatings on mineral sands are lightly-adhering clays which can be removed quite readily by agitating the sands in water for a few minutes-thus their removal constitutesno problem. However, the heavy gingery coloured coating which adheres strongly to the surface of the particles in some deposits, and which has generally been considered to be a form of iron oxide (Baker, 1962; Hudson, 1966), cannot be removed effectively by attrition cleaning in water (Hudson, 1966). Such coatings limit the usefulness of these concentrates in the foundry industry.An investigation of strongly-adhering coatings on "ginger zircon" concentrates from the Yoganup deposit of Westralian Sands Ltd has shown that the stain consists of a cemented mixtvre of hydrated iron and aluminium oxides and kaolin minerals, and that it can be removed by a relatively low cost treatment.IDENTIFICATION OF THE STAIN BY XRDX-ray powder diffraction (XRD) patterns were obtained from small samples of stain scraped, individually, from the surface of hand-picked zircon and ilmenite crystals. These scrapings were mounted on quartz fibres and exposed to CoKa radiation in a 57 mm Debye camera. To avoid heavy fogging of the x-ray films, a substantial amount of the iron oxides was preferentially leached from the scrapings (before mounting) with an ammonium oxalate-oxalic acid solution (Rooksby, 1961). The residual material gave a fairly well resolved powder pattern (Table 1). A...

Jan 1, 1975

By G Newling

Zinifex Rosebery Mine is a lead-zinc underground mining operation located at the base of Mount Black, on the West Coast of Tasmania. Mining activities at the Rosebery Mine have been undertaken for over 100 years, and it has been operating in its current form for over 60 years. A large surface and underground exploration drilling program over the last 18 months has increased the mineÆs expected mine life from five to 12 years, with expectations that over the next 18 months as the program continues this will swell to 20 plus years. Zinifex Ltd secured Rosebery Mine in April 2004 and commissioned a detailed life-of-mine (LOM) study on all aspects of the operation. One of the most detailed sections of this report was on the state of the current ventilation system. It was identified in this report that for mining at the current depth of approximately 1100 m to continue a minimum 30 per cent increase in ventilation quantity was essential, as well as a likely requirement for a refrigerant cooling plant for each of the two deepest panels, the K and P lenses. It was identified that Zinifex did not have personnel with the expertise in underground refrigeration to manage and implement a project of this nature. A second report was commissioned with Bluhm Burton Engineering (BBE), a South African firm specialising in both surface and underground large-scale refrigeration plants. BBE concluded that two refrigeration plants were required to provide the necessary cooling power in the lower sections of the mine. In 2005, with BBE as technical consultants, Zinifex Rosebery Mine embarked on a project to commission two underground refrigeration units capable of lowering the wet bulb temperature at the decline face by four degrees centigrade. This paper is intended to give the reader an overview of: Rosebery Mine operations,Rosebery underground refrigeration project,technical difficulties associated with implementing an underground refrigeration system, andkey learnings to date.

Jan 1, 2008

By X Zheng, P Manton

Pebble recycle crushing circuit performance at Telfer has historically fallen significantly short of the design product P80 of 12 mm. An investigation was undertaken to evaluate the current pebble crusher circuit performance, quantify the optimum performance achievable with the existing circuit and examine the impact of pebble crusher performance on the overall grinding circuit throughput. The primary objective of this investigation was to develop an operating strategy that could consistently deliver the finest possible pebble crusher product to maximise the overall plant throughput.The pebble crusher specific energy (Wc) was proposed as a key control variable for pebble crusher operation based on the results of plant survey work. Over the range of pebble crusher specific energies encountered at Telfer, a strong linear relationship was found to exist between Wc and the crusher product size. This direct linear relationship between specific energy and crusher product size remained constant regardless of how the change in specific energy was generated, ie through changes to the crusher closed side setting (CSS), feed rate, liner profile or moisture content. Key findings for the Telfer pebble crushing circuit were: each 0.1 kWh/t change in Wc resulted in a 4.9 per cent to 6.0 per cent change in plant throughput, each 1 mm change in pebble crusher CSS resulted in a 2.2 per cent to 2.6 per cent change in plant throughput, and 0.57 tonne of new feed was displaced for each tonne of crushed pebbles recycled to the semi-autogenous grinding (SAG) mill. The presence of an existing 3000 tonne pebble stockpile between the SAG mill and pebble crushers provides an opportunity to manipulate the feed rate to the pebble crushers independently of the SAG mill pebble production rate. This flexibility, along with excess pebble crushing capacity, enables the pebble crusher specific energy to be maximised the majority of the time. Pebble crusher specific energy has subsequently been incorporated as an online crusher control measure to assist in maximising plant throughput.

Jan 1, 2009

In mineral asset valuation the use of any valuation method requires the input of competent and independent technical professionals. A resource valuation must be as transparent as possible to allow others to assess the validity of their underlying assumptions. Materiality is a dominant factor in all valuations and a valuer should set out the basis for the subjective decisions made in this regard. Any valuation methodology must comply with the requirements of the law, practice notes and guidelines issued by the various relevant authorities and professional bodies. The valuation methodology chosen to value a mineral asset depends upon the amount of data available on that asset and the reliance that can be placed upon those data. Mineral assets generally fall into one of three broad categories, depending upon their state of development, ie whether they are at the exploration, development or operating stage. Naturally there are transition phases. Hence, there are ""Grass Roots"" Exploration Areas and Advanced Exploration Prospects; Pre-Development Projects and Developing Mines; and Operating Mines. The features of each of these mineral asset status categories are described, mainly in terms of the confidence in the contained identified Resources/ Reserves. The various subjective methodologies available for the valuation of ""Grass Roots"" Exploration Areas and Advanced Exploration Prospects -are specifically reviewed. The pitfalls in adopting a ""recipe"" approach are noted. It is the author's view that the Expected Value Method (based on the probability adjustment of the output of a Discounted Cash Flow/ Net Present Value model) has little application in the valuation of exploration mineral assets, so it is only briefly addressed. The methodologies reviewed are listed below. ò Multiple of Exploration Expenditure Method (expenditure based). ò Joint Venture Terms Method (expenditure based). ò Geoscience Rating Methods (eg Kilburn - area based). ò Comparable Market Value Method (real estate based). ò Rules of Thumb (eg $/Resource or production unit; % of in situ value). Some recent Prospectuses are used to provide examples of their misinterpretation and misapplication.

Jan 1, 1994

By N S. Rauert

The Barrick Gold Granny Smith underground operation located at the Wallaby deposit is currently in a transition period from a near surface operation in ramp up mode to a fully operational deep mining operation with annual production exceeding 1 Mt/a. A key element in this process has been the project execution of the Wallaby Deeps project from initial scoping study in 2007 to the major production component of the Wallaby operation by 2013.The Wallaby Deeps project serves as an example of how an ongoing detailed study program can be integrated into a day to day mining operation assessing risk and mining application in a perceived challenging narrow vein geotechnical environment prior to mining. A detailed study phase to feasibility level is carried out prior to the mining each zone of Wallaby with mining approval only granted on study completion with full justification.A central element in this process is an optimisation study that utilising the Mineable Shape Optimiser (MSO) and Schedule Optimisation Tool (SOT) software tools developed by AMIRA Ltd PRIMO research project is used to carry out an NPV value based option study. This optimisation process being then tempered to reality by a predefined practicality assessment process utilising extensive peer review by operational and study team members as well as industry experts.Other challenges facing the Wallaby Deeps project include the requirement to maintain the ramp up production profile to full production by 2012 without disruption but with pressures of an ongoing geological resource to reserve conversion program essential for the study process and mining approvals and also development rate achievement.To date these challenges are being met with integrated planning and the value of the study program itself. Production in 2012 year to date is well on the way to annual production rate exceeding 1 Mt/a.CITATION:Rauert, N S, 2012. Wallaby Deeps - from the optimal opportunity to reality, in Proceedings Narrow Vein Mining 2012 , pp 207-214 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Mar 26, 2012

By S Januszewski and, G Dirom, D Christensen

Four field test plots were constructed in 1998 on a sloped waste rock surface at the Myra Falls Operations of Boliden Westmin (Canada) Limited to evaluate hydraulic performance of alternate cover system designs. The field test plot cover systems were designed to control oxygen ingress and water infiltration to the underlying waste rock while providing a medium for a grass and legume vegetation cover. The test plots consisted of a bare waste rock surface (control plot), and three additional plots consisting of native compacted till placed directly on the waste rock surface, overlain by non-compacted till. Two of the test plots included a compacted till layer ameliorated with flyash and bentonite, respectively, to improve moisture retention and reduce permeability. The third test plot consisted of a compacted layer of native till overlain by non-compacted till. A performance monitoring system was installed in each test plot to evaluate net percolation and oxygen ingress to the underlying waste rock. In addition, each test plot has profiles of matric suction, water content, and temperature sensors installed at up-slope and down slope locations. A review of four years of monitoring data shows that the compacted till/bentonite cover system has been most successful in limiting the infiltration of meteoric waters and oxygen ingress to the underlying waste rock material. The compacted till/flyash and compacted till cover systems have produced similar net percolation results and neither have functioned well in maintaining an oxygen ingress layer throughout the year. The amelioration of bentonite with till to produce a compacted barrier layer shows promise in reducing water and oxygen ingress at the MFO site, while the incorporation of flyash with the till did not result in a significant increase in performance. The field performance data clearly demonstrates the need for developing field trials prior to full-scale construction, as opposed to relying solely on an un-calibrated numerical modelling to design the full-scale cover system.

Jan 1, 2003

The authors have succeeded in providing additional rigour and clarity to the formulation and execution of the dynamic programming algorithm for optimum open pit limit determination. The inclusion of the blocks on the J + 1 column in the search for Pmax, and constraining this search with blocks assigned a value of negative infinity, ensures an optimal solution within the defmed blocks. The effect of the negative infmity blocks is to preclude mining along the diagonal blocks from the first row that they occur in. Figure 1 graphically illustrates the effects of this on the calculation of P. A practical application for constraining mining with negative infmity blocks is the case where the pit outline is defined by lease boundaries and not by the value of blocks to be removed. Any legal, social or physical constraint could be accounted for in the same way.The danger in applying this technique is that the negative infinity blocks will override any other more profitable solutions that may exist. This is in fact what has happened in the example that the authors use. If a further column is added to the left hand side (column -1) and the calculation is carried out again, a more optimal x-section can be found. (If the negative infinity values are removed the same result is obtained.) This gives a pit value of 112 as against the 108 of the authors' optimal pit. This calculation is shown in Figure 2. Lerchs-Grossman optimization is reliable only if there are sufficient columns available to allow "mining" to optimum depth of the ore body and still allow the pit to break out at the surface within the confines of the columns.This author's approach is to create enough columns each side of the lateral limits of the ore to allow the pit to intersect the surface while still within the confines of the block model. Implementing the methods discussed by the authors helps to reduce the number of columns required to achieve this.

Jan 1, 1987

The carbonatite-hosted apatite deposit at the Cajati open pit mine in the Jacupiranga igneous-carbonatite complex is located in south-eastern Brazil. The major styles of apatite mineralisation include veins or veinlets, small pods or patches, disseminated apatite and locally massive aggregates. The metallurgical concentration of the apatite is accomplished by grinding the carbonatite, removal of magnetite with a magnetic drum, size classification with cyclones and flotation to suppress carbonates and collect apatite. To examine the character of the apatite grains and to apply those observations to apatite beneficiation, specimens of carbonatite were collected from recent exposures in the mine and from selected drill holes. This study utilised a variety of microscopy methods, including transmitted and reflected light, cathodoluminescence (CL) and scanning electron microscopy with wavelength-dispersive spectrometers for mineral compositions. It was found that LREE, Na and Sr are the main elements replacing Ca in the apatite, with relatively low oxide contents (<0.5 per cent, but up to about one per cent for Sr). LREE and Sr contents consistently increase in the latest intrusions (C4, C5 and SP24). Chemical and morphological modifications of apatite were recognised to be related to detrimental effects on flotation recovery. Three examples illustrate this relationship: 1.  Apatite with reddish CL is fractured. Locally that apatite exhibits no CL due to CL quenching caused by iron entering the apatite structure.2.  Apatite from the xenolith-rich area exhibits abundant fine, platy silicates that appear to interfere with flotation and lower apatite recovery.3.  So called æhardÆ apatite, which is more difficult to grind, occurs in an area in the mine that was affected by hydrothermal activity. That activity is evidenced by the development of a bright, white CL rim in the apatite, and by the presence of late veins of carbonates and sulfides that fill fractures and appear to contribute to higher cohesiveness of those apatite grains.

Jan 1, 2008

By P N. Holtham

Magnetic separation is routinely used in the processing of titanium minerals, and the efficiency of these operations can be determined by measuring the magnetic properties of the process streams. The valuable components in the feed to a titanium minerals plant have specific magnetic susceptibilities ranging from -10-9 m3 kg-1 (zircon) to 10-6 m3 kg-1 (ilmenite). However, in an industrial environment, in which the mineral temperature may range from 20 to 120¦C, it is difficult to measure susceptibilities lower than about 10-7 m3 kg-1. Hence laboratory testing of spot samples from magnetic separators is currently required to obtain accurate performance data. This is an inherently slow process and where feed grades are variable, optimum performance of magnetic separators is generally not achieved. This paper describes the development of an inductance-based instrument for measuring very low levels of magnetic susceptibility on-line. The signal from the instrument can be used for process control purposes to maintain optimum magnetic separator performance. A prototype instrument has been installed in the zircon scavenger circuit in an Australian titanium minerals plant. The circuit consists of a roll magnetic separator treating a feed consisting mainly of monazite and stained zircon. The monazite-rich magnetic fraction is returned to the mine site for disposal, while the non-magnetic fraction is recycled to the zircon wet circuit. Splitter positions in the separator are adjusted when the final zircon product fails to meet specification, or the magnetic fraction contains more than 30 per cent zircon. Weekly composite grain counts show that sub-optimal operation of the magnetic separator results in significant zircon losses into the magnetics fraction. Using the instrument, the zircon content of this stream can now be accurately determined from the measured magnetic susceptibility, provided a correction for the effect of mineral temperature is applied using the Curie - Weiss relationship.

Jan 1, 2001

The silver-lead ores of Yerranderie occur in a series of narrow, flat-dipping lodes in small bodies of quartz porphyry, intrusive into Upper Devonian volcanics, which are exposed in the western flank of the Permo-Triassic Basin of central New South Wales, about 80 miles south-west of Sydney (Harper, 1930; Lawrence, 1953). The lodes strike about E.-W. and are from 1,200ft. to 2,000ft. long. They consist of a series of ore shoots from 3ft. to 100ft. long, and up to 5ft. thick, though commonly under 2ft. thick, interspaced with low grade sections. They dip at from 18°S. to 400 S., and have persisted down the dip for 1,450ft. (693ft. vertical depth) in the Wollondilly mine, for 1,120ft. in the Bartlett's mine, and for 850ft. in the Wonga mine.The grade of ore in the Bartlett's mine was from 32% to 40% lead, from 100 oz. to 130 oz. silver per ton, 0.5 to 4.0% arsenic, with ore ,as rich as 20% arsenic on the No. 6 and No. 7 levels of the Silver Peak mine, the underlay extension of the Bartlett's lode, about 1% copper, and up to 4 dwt. gold per ton. In the Wollondilly mine the average grade was 15% lead, with sections as rich" as 36% lead, from 75 to 165 oz. silver per ton, 0.6 to 3.3% arsenic, 0.2 to 0.7% antimony, 0.8% copper, and up to 7...5. Cl~t. gold per ton. The zinc content of the ores is slight, and has not been recorded.Mining began about 1898, and to date more than 110,000 tons of ore, valued at £1,800,000, has been raised. Large scale mining ceased about 1935. The chief mines operating were the Bartlett's mine (formerly Colon Peak), the Silver Peaks, which is the underlay extension of the Bartlett's, the Wollondilly (or Yerranderie) mine, and the Wonga' (or Tonalli) mine.

Jan 1, 1953

Over the past seven years, a combination of New Zealand-, Australian- and German-led research voyages using the research vessels Tangaroa, Southern Surveyor, and Sonne respectively, have revolutionised our understanding of the Kermadec-Tonga arc-backarc systems. This convergent margin is the most seismically-active subduction zone system on Earth with convergence rates steadily increasing northwards from the Hikurangi Trench to ~250 mm/year at northern Tonga. In combination with the intra-oceanic setting, factors potentially producing magmatic, tectonic, and hydrothermal variability in this system include an increase in potentially subductable sediment northwards along-strike, intersection of the Louisville Ridge (hot-spot chain) and Osbourn Trough (paleo-spreading centre) with the arc at ~25¦S, and the possibility of Samoan (hot-spot) plume ingression in the northern Lau Basin. The bulk of the magmatism is submarine; whereas only a few emergent (some transient) volcanic islands exist, about 80 submarine volcanic centres of significant size (>1000 m above surrounding sea floor) are present between White Island in the south to north of Tafahi in Tonga. About 35 per cent of these centres are hydrothermally active. Important petrogenetic features of the volcanic rocks along this arc system are the abundance of rhyolite (some quartz-amphibole-bearing), and the outcrop of primitive olivine-clinopyroxene-phyric basalts in satellite cones of the major edifices. Many of the edifices consist of complex caldera systems. Eruptions from and degradation of these edifices were probably tsunamigenic. Many of the hydrothermal systems developed within the edifices are likely destroyed during the sector collapses which are the characteristic edifice degradation mode. There are three backarc spreading systems along this margin: the Havre Trough, Valu Fa-Eastern Lau Spreading Centre, and the nascent Fonualei Rifts. The latter are particularly notable in having captured the entire suprasubduction zone magmatic flux for a distance of about 150 km, shutting down the adjacent volcanic front arc volcanoes.

Jan 1, 2005

By Abbott T. M

Alcoa of Australia operates three Alumina refineries in the South West of Western Australia at Kwinana adjacent to the South suburbs of Perth and at Pinjarra and Wagerup in rural settings 100 and 150 kilometres south of Perth respectively (see Figure 1). Alcoa's total alumina production is currently around five million tonnes per annum and this will increase to 5.5 million tonnes by early 1990. Further expansion of production is likely if current world aluminium demand trends continue. Alcoa's refineries utilise bauxite mined in the nearby Darling Range which is low grade by world standards having around 32 % by weight extractable alumina (by comparison Weipa bauxite yields around 55% alumina). This means that for every tonne of alumina produced two tonnes of solid wastes remain and therefore while we in Western Australia can boast of hosting the largest alumina industry in the world with around 18% of total world production we suffer from having the lion's share or 40% of the world's bauxite residues to dispose of. In its Western Australian setting the management of these residues in a manner which is acceptable to the community creates many challenges. These are primarily due to the mass and volume of waste produced, its chemical and physical characteristics and the environmental features of the local area including close proximity of neighbours, competing land uses and the presence of natural resources such as surface water and groundwater. In addition the trend towards more stringent statutory environmental controls and greater political and community awareness of environmental issues means that we must be able to demonstrate improvements that will reduce the considerable adverse impacts of past practices. Of course we must also try to keep the overall cost of residue disposal as low as possible to maintain our competitiveness with refineries which may have much smaller volumes of waste to dispose of and less environmental constraints.

Jan 1, 1989

By Heyer T, Wasmuth H-D

The concentrator plant of Fahrica Mine, that started production in 1977, was using nine Jones wet high intensity magnetic separators (WHIMS) of the standard type DP 317 with a 1.5 mm gap width for the production of pelletfeed concentrate. These WHIMS units were operating in parallel with a feed rate of 110 t per hour, with about 47 per cent Fe in the feed, and with a maximum of only 0.8 per cent magnetite. The field strength was about 1.2 Tesla. From 1986, the mine has adapted its production facilities to mine and concentrate larger amounts of lowgrade itabintes to maximise the use of the available iron ore reserves. An increase of up to three per cent magnetite in the magnetic separator feed, resulted in a reduction of capacity. This was due to partial clogging of the separation boxes at high field intensities, which were required for the recovery of the very fine hematite. Even by lowering the feed rate and the field intensity the problem could not be resolved. After extensive testwork in a pilot plant, with a Jones separator model DI' 90, the commercial WHIMS plant was modified into a two stage process using the existing nine Jones separators. The gap of the first stage with four machines was increased to a width of 3.5 mm, corresponding to a maximum field intensity of 0.7 Tesla. This first 'medium intensity scalping stage' is followed by a second 'real' high intensity WHIMS step with five machines operating at the original settings with 1.5 mm gap width and 1.0 Testa maximum field intensity. Feed rates of up to 320 tons per hour in the first and up to 240 tons per hour in the second stage could be achieved. The design concept of the concentrator at Fabrica mine is described using the old and new flowsheets. Details of the pilot plant testwork as well as actual performance data of the new plant section are presented.

Jan 1, 1993

Oceana Gold CorporationÆs (OGC) Frasers underground (FRUG) transitioned into production in early 2008. The mine uses a relatively unique yielding pillars mining method that takes its lessons from empirical coal mining practices more so than hard rock techniques. Being a low-grade, flat-dipping resource hosted in the highly structured relatively weak Otago schist in central South Island in New Zealand, meant a number of iterations were undertaken to find the right mix of mining cost and resource extraction. Mining is on the down dip extension of the active Frasers open pit at the Macraes site (that extracts the full shear zone package at a head grade of 1.5 g/t). The challenges for the FRUG feasibility study team was to deliver an net present value (NPV) positive operation that would exist as a coproduct with OGCÆs other New Zealand operations, allowing it to make use of the high throughput low cost refractory ore treatment facility.Key factors in the successful project design were: Net capital costs kept to less than NZ$30 M by realising significant gold revenues from mine development from month four onwards. Virtually eliminating waste development by putting all key capital infrastructure in the low-grade ore adjacent to panels and maintaining the option of extracting this resource on retreat from the mine. Targeting a specific higher grade halo ore package within the entire mineralised shear zone. Applying a mining method that accepted a higher degree of dilution, ore loss and resource loss (in yielding and non-yielding pillars) to the flat-dipping resource. Thus providing an acceptable margin per tonne and resource tonnes mix, as the resource could not support a conventional low-cost caving operation, nor a higher cost room and pillar or open stope and backfill operation. 2008 has seen increasing tonnes and grade performance with a corresponding decrease in cost per tonne and cost per ounce as the mine has reached nominal capacity and is producing an operating surplus. The challenge will be to maintain and improve.

Jan 1, 2009

By M Morey

Ok Tedi Mine is located eighteen kilometers from the Irian Jaya border in the Star Mountains, Western Province, Papua New Guinea. Sequential flotation concentrators produce a copper / gold concentrate for market, and a pyrite concentrate for environmental containment, with low sulphur tailings from the pyrite flotation sequence being discharged to river. Strict sulphur assay targets are set for the riverine discharge. Current porphyry and skarn ore blends from open pit operations at Mt. Fubilan will cease by 2017. However, drill cores indicate an economic ore body in the skarn ores surrounding and below the current pit limits. With proposed mining from a western pit wall ‘cut back’ and an eastern pit wall underground operation, potential for ore ageing in sub level cave underground stockpiles is apparent. With little information in the literature, Ore Ageing test work at prefeasibility level consisted of multiple one tonne stockpiles of -50 mm crushed rock. For the bankable feasibility study, a successful test work program was implemented utilizing 40 tonne ROM ore stockpiles composed of exposed pit extensions of future ores. Realistic ageing profiles of Skarn ores were obtained from one tonne monthly samples subjected to laboratory testing. Test work outcomes included:• Similar ore ageing profiles from both the pre feasibility study (PFS) and feasibility study (FS) programs, suggesting ore susceptibility to oxidation has greater effect in ore ageing than rock size for Ok Tedi skarns. • Laboratory rougher flotation of pyrite showed significant variation in sulphur recovery rate, suggesting three classes of pyrite floatability.• Laboratory rougher flotation of both copper and pyrite for up to 19 months ore ageing from 11 stockpiles, suggested three stages in the ore ageing process.

Aug 8, 2011

By P Ridley

EnergyAustralia’s Sydney Central Business District (CBD) cable tunnel network will form a loop 6.2 km in length. The 3.2 km long, 3.5 m, diameter city east cable tunnel (CECT) currently in design, will complete the loop in 2014. This paper discusses the ventilation design of the CECT, considering operation for four functions:1. normal (unmanned),2. access (manned),3. purge, and4. emergency fire modes.The dominant design criterion which determines the required ventilation capacity for CECT is the specification of an upper air temperature limits of 35°C (dry bulb) and 24°C (wet bulb) for manned access, during all ambient weather conditions and cable heating loads. Simplistic static heat transfer analyses do not yield a sufficiently accurate calculation of tunnel air temperature and humidity to provide a safe and yet economical design. The primary refinement in the analysis discussed in this paper is associated with modelling of the system’s dynamic behaviour.Our means of analysis is through a thermo-fluid-dynamic simulation of the interactions between the diurnally and seasonally varying loads (cable heat generation, ambient air conditions and groundwater inflow), tunnel airflow and the resulting tunnel air temperature and humidity. These simulations are validated using experimental data from the existing city south and city west cable tunnels.Fire life safety scenarios are also analysed and discussed in association with the ventilation design. Ventilation strategies are developed for dealing with such emergency fire situations.The understanding and insight provided by the work offers the ability to optimise ventilation plant provisions in future cable tunnels, particularly long and high energy density tunnels.

Mar 8, 2011

By B Roache, A Vakili, R Talebi

When a highly anisotropic or weak rock mass is subjected to high-stress conditions, it can lead to large ongoing deformations or squeezing ground conditions. Squeezing ground conditions can impose increased support costs, delays associated with rehabilitation, and disruption to production schedules. Ultimately, in more severe cases, it may result in complete drive closure. Several recent case studies investigated the squeezing mechanism and measures to control squeezing ground in underground mines and also to design appropriate ground support systems. Effective management of squeezing ground conditions requires a sound understanding of the failure mechanism and the driving forces. Once the failure mechanism is understood, the appropriate ground support system should be tailored to specifically manage the critical driving forces that control the squeezing mechanism. Support type, support stiffness, reinforcement length, support density, installation times and mining sequence all have important impact on the effectiveness of the support system. This paper provides a case study at the Ballarat Gold Mine in Australia, where drive squeezing was associated with foliated ground conditions located below 700 m depth. Numerical modelling was used to better understand the failure mechanism and the driving forces that lead to squeezing conditions. The effectiveness of various support types was compared to installed ground support performance. This has provided a credible methodology for assessing and optimising the selection of reinforcement in squeezing ground conditions. CITATION:Talebi, R, Roache, B and Vakili, A, 2018. Understanding and modelling squeezing ground conditions at the Ballarat Gold Mine, in Proceedings The Fourth Australasian Ground Control in Mining Conference (AusRock), pp 107–118 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 30, 2018

By P K. Seccombe

The McKinnons gold deposit is hosted by weakly deformed, Palaeozoic metasedimentary rocks of the Cobar Basin in the western Lachlan Fold Belt of New South Wales. Mineralisation comprised a largely supergene resource of 1.7 Mt of ore, grading 1.7 g/t gold, established over structurally controlled, primary gold and base-metal mineralisation in underlying, carbonate-rich meta-siltstones and sandstones of Early Devonian age. Mineralisation was associated with at least two major fluid cycles. The first cycle was characterised by four stages of vein formation. Early quartz and carbonate veins (Stages 1 and 2) and quartz-pyrite breccias (Stage-3) occurred prior to, or early in, the main regional deformation (D1). Continued D1 deformation produced syntectonic, carbonate-base-metal veins (Stage-4), primarily in dilatant zones adjacent to fold axes. The second cycle of mineralisation was an epithermal-style event, driven by a shallow, syn-D2 or post-tectonic intrusion, defined by aeromagnetic imaging. The intrusion may have exploited a jog, controlled along the basin margin by basement faults. Mineralisation produced gold-pyrite veins, associated with pervasive quartz-sericite-adularia and chlorite-sericite alteration. Highest primary gold grades are confined to quartz-adularia boiling zones. The McKinnons deposit offers an important example of mineralisation formed by multiple processes. Early-stage, base-metal mineralisation at McKinnons reflects structurally focussed, basin dewatering and metamorphic processes, typical of other deposits in the Cobar Basin. By contrast, late-stage gold mineralisation at McKinnons is magmatically driven and of epithermal style. Recognition of overprinting distinctive geological events is critical, not only in modelling ore formation, but in developing a versatile exploration concept.

Jan 1, 1999

By Y K. Ji, B J. Chung

In recent years, the utilisation of ultra-fine ore in ironmaking has gradually increased due to the depletion of high-quality iron ore. The use of ultra-fine ore in the ironmaking process is typically energy intensive and leads to high CO2 emissions. The improvement of sinter reducibility is one of the methods used to reduce the environmental impact of using fine ore.In this study, a high-strength and high-reducibility sinter manufacturing process was investigated by controlling the sintering reaction. The key technology used in this method is the manufacture of mini-pellets by selective granulation equipment. The mini-pellets consist of iron ore, limestone and burnt lime and are designed to melt at temperatures below 1200°C. The molar ratio of CaO to Fe2O3 of the mini-pellets is 0.25~0.75. The 2~5 mm diameter mini-pellets are manufactured with burnt lime as the binding material.During the sintering of mini-pellets, goethite, which has relatively high reducibility, should be retained as residual ore in the sinter cake. The strength of the sinter cake was kept by the bonding reaction of the mini-pellets, which have a low melting temperature. This is the main principle of the reaction-controlled sintering method.The reaction-controlled sintering method manufactures a high-strength and high-reducibility sinter cake and increases sinter productivity by around 15 per cent using ultra-fine iron ore.CITATION:Chung, B J and Ji, Y K, 2015. Production of high-quality sinter by controlling the sintering reaction, in Proceedings Iron Ore 2015, pp 21–28 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 13, 2015

By J Theiler, B Beamish

The presence of reactive metal sulfide minerals in waste rock contributes to the hazard of spontaneous combustion in dumps, as these minerals react with air and water entering the dump at ambient temperatures to form various mineral products and significant heat. This reaction series can also be an initiator of self-heating of carbonaceous material present in the waste rocks if sufficiently high temperatures are reached. Characterisation of these reactive minerals has primarily been to assess the risk of acid mine drainage, which is not necessarily an indicator of spontaneous combustion risk. No systematic testing procedure has been adopted by industry to assess the spontaneous combustion propensity of the waste rock. As a consequence, the principal management strategy for spontaneous combustion has been a reactive approach of dealing with events as they arise.Recent advances in adiabatic oven-testing have enabled more definitive assessment of the spontaneous combustion propensity of coals, including the quantification of the accelerating influence of reactive pyrite (Beamish and Beamish, 2011, 2012; Beamish, Lin and Beamish, 2012). This same test procedure has been used to assess the self-heating performance of metal concentrates and is therefore amenable to assessing the metal sulfide oxidation process from low ambient temperatures. The data obtained from such testing can be used to quantify the more prone waste rock sequences of a mineral deposit in advance of mining. This enables appropriate management planning for emplacement and long-term storage of waste rock in dumps.CITATION:Beamish, B and Theiler, J, 2016. Characterising the spontaneous combustion propensity of waste rock, in Proceedings Life-of-Mine 2016 Conference , pp 93–96 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jun 28, 2016