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By K Ehrig, V Liebezeit, Y Li, B Pewkliang, M Smith, E Macmillan

Pre-concentration of low grade ores via the rejection of low/no value materials based on fragment size is the subject of many published papers and studies over the past decade. However, the possibility of upgrading should not limited to low grade ores given the high cost of capital, energy, water, reagents, and ultimately tailings storage across the mining industry. Mining and processing facilities at Olympic Dam consist of an: 1) underground mine where ore is extracted via sublevel open stoping and 2) integrated processing plant located on the surface which produces Cu-cathode, Au-Ag bullion and uranium oxide concentrate. Blasted ore is trucked/trained to underground crushers which produce an ore feed with a product size P80 = -145 mm. Crushed ore is then hoisted to the surface and delivered to the ROM pads via an overland conveyor. The typical copper grade of mill feed is approximately 2 wt%. A rapid assessment of the potential to upgrade Olympic Dam mill feed prior to the ROM pads via rejection of low value materials based on fragment size was completed using existing data. Three separate datasets were utilised to assess the deportment of Cu, U, Au across size ranges of 200 mm down to sub-8 m. Cumulative grade vs cumulative mass distribution graphs for each element were produced. This data clearly demonstrates that Cu, U, Au grades remain constant down to ~212 m, which happens to be the liberation size of the sulfides. Hence, no potential to upgrade the ore stream via rejection of +200 mm - to ~212 m sized materials prior to milling. Mineralisation within the Olympic Dam Fe-oxide Cu-U-Au-Ag deposit (IOCG) occurs as disseminated sulfide grains within granite- to hematite-rich breccias. The typical sulfide grain size is sub-200 m that supports the outcomes of the grade deportment by fragment size study.

Nov 10, 2020

By A E. Annels, G F. Johansen, B W. Cuffley

Nuggety gold-quartz reefs are characterised by localised, high-grades and in many cases coarse free-gold. Geometrical and grade complexities are often observed. The high-grade regions are generally erratic and have a low continuity; though may make the deposit economic. Viable grades are generally contained within discrete oreshoots, which are surrounded by barren/low-grade material. Drilling is an effective measure of geological continuity, however grade distribution can only be reliably obtained from underground development and bulk sampling. In general terms, only an inferred resource is likely to be estimated from surface drill data alone. Underground development, further drilling and probably bulk sampling will be required to delineate inferred and indicated resources. Closely spaced development with bulk sampling is likely to be the best way to determine a probable reserve. The resource evaluation process must driven by a clear geological model that should attempt to understand the controls to reef geometry and grade distribution. Geological interpretation has traditionally been undertaken in two-dimensions, though computer software is now being used to produce three-dimensional models. Grade interpolation has generally been undertaken using classical methods, typically weighted averages, zone of influence weighting and polygonal/sectional blocks. Tonnages can generally be calculated from diamond drill and development information with a reasonable degree of confidence. Grade is much more difficult to define with confidence because of its highly erratic and discontinuous nature. To account for variability it is recommended that a best estimate be reported within a grade range, as this more accurately reflects the uncertainties involved. Reporting should be undertaken within the framework of the JORC code, but should involve special consideration of grade reliability.

Jan 1, 2000

By M Astorga, J Jackson, J Gaunt

The Ban Houayxai mine (BHX) is a relatively low-grade, low-cost, open pit gold-silver deposit in Laos operated by Phu Bia Mining, a subsidiary of PanAust. The ore production rate at BHX is 4.5 Mt/a, with direct tipping to a semi-autogenous grinding/ball mill circuit with a carbon-in-leach process plant. Approximately 100 000 oz of gold is produced per annum.The operation is located in mountainous terrain with minimal run-of-mine stockpiling. This results in limited capacity for blending from stockpile, with the mill instead reliant upon direct feed ex-pit. Since commissioning in 2012, the plant has seen significant variation in milling rates due to variability in the feed properties of the oxide, transition and primary ores. As part of an ongoing continuous improvement program, an integrated approach was initiated, focusing on maintaining and enhancing production in the future as the proportion of harder primary ore increases. This will specifically focus on direct blending from the loading face using the ore properties and blast fragmentation to maintain mill throughput.This approach was based on the concept of physical assets management, commencing with improving information and knowledge of the condition of the orebody through modelling the characteristics, variability and performance of the feed for processes relevant to throughput. These models are used to support both mine and process plant production planning.The key ore feed characteristics and parameters modelled for the life-of-mine are blastability index, powder factor, crushability or impact resistance and grindability. These predictive spatial models were based on the data from diamond drill holes used in resource definition and geotechnical drilling programs by integrating geotechnical, geological, geochemical and metallurgical data.At the early stages of implementation, the models are being utilised for ore blending decisions to provide guidance and support for budgeting, long-term mine planning, blast design for mill feed and providing the mill with an expectation of performance.CITATION:Jackson, J, Gaunt, J and Astorga, M, 2014. Predicting mill ore feed variability using integrated geotechnical/geometallurgical models, in Proceedings Orebody Modelling and Strategic Mine Planning Symposium 2014, pp 165–176 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 24, 2014

By L R. Sousa, M He

Rock burst is common in deep underground excavations and is characterised by a violent ejection of block rocks from excavation walls. It is critical to understand the phenomenon of rock burst and particularly the failure mechanism. Laboratory experiments are one of the ways. For that purpose, a rock burst laboratory test system was developed at the State Key Laboratory for Geomechanics and Deep Underground Engineering at the China University of Mining and Technology, Beijing. A true triaxial test system, designated the Deep Underground Rock Burst Analogue Test Machine, was developed, which can transfer a stress state from three-direction-compression on six surfaces to the compression on five surfaces with one set free. The rock burst experimental system comprises an acquisition of acoustic emission signals, a high-speed recording in order to accurately record the kinetic characteristics of rock fragments ejected during a rock burst event and the possibility of infrared thermography showing the surface temperature of the samples. In this paper, analyses of typical results from rock burst tests are presented with a description of the obtained results. The mechanisms of rock burst are illustrated, and a rock burst classification is proposed based on these laboratory tests. For the control of rock burst, the constant resistance, large deformation (CRLD) bolt or anchor is presented for accommodating the large displacement of surrounding rock masses and absorbing the impact of the sudden release of rock burst while outputting a constant resistant force in response to the external load. The CRLD bolt or anchor has had practical applications in underground mines. Finally, in situ tests performed in a tunnel at Qingshui coalmine, China, are analysed in detail.CITATION:He, M and Sousa, L R, 2014. Experiments on rock burst and its control, in Proceedings AusRock 2014: Third Australasian Ground Control in Mining Conference , pp 19–32 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 5, 2014

By A G. Sheard

Fire safety requirements inside tunnels have become more demanding as tunnel designers consider the impact of potentially larger fires. Tunnel ventilation equipment is required to clear hot gases in the event of a fire. This is vital to maintain smoke free escape routes for people trapped in the tunnel and to facilitate safe access for emergency services. As the fire becomes larger, the temperature of clearing hot gases increases. Traditionally, engineers have taken into consideration gas temperatures of 250°C when designing tunnels. Because of larger fires, engineers must now design equipment to clear hot gas up to 400°C.Increasing the temperature of hot gases to 400°C requires tunnel ventilation equipment that goes beyond historic limits. As temperature rises, material strength reduces, consequently mechanically constraining fan aerodynamic performance. Developing tunnel ventilation equipment suitable for clearing hot gases at 400°C requires a fan technology that is more typical of industrial turbomachinery. The standards EN 12101-3:2002 and ISO 21927-3:2006 defi ne the necessary design criteria and testing methodologies to verify that a range of tunnel ventilation equipment is capable of extracting hot gases.This paper describes the process by which the authors certified in accordance with EN 12101-3 and International Organization for Standardization (ISO) 21927-3 a range of powered ventilators suitable for clearing smoke and hot gases at 400°C for two hours. Proving the high temperature capability of the new range provided a challenge as a consequence of the large physical sizes and high motor powers involved. Additionally, the need to install vertically, as well as horizontally, imposed high thrust loads on motor bearings, further complicating both mechanical design and testing. The paper presents the methods by which the authors overcame these challenges.

Mar 8, 2011

By G J. Halstead, R C. Steinhaus

The True Pipe® in-line sampler was developed in an attempt to address the need for a mechanically reliable and correctly designed sampler to suit one-dimensional pressurised slurry pipelines, often found in metallurgical process plants but for which no reliable and representative sampling solution as yet exists.The design, whose operation is based on the principle of symmetry, is described by Dominique François-Bongarçon (2005) with the synchronous opening and closing of relevant valves. It representatively isolates portions of the particulate material stream at predetermined intervals for effective extraction of primary increments for subsequent sample compositing.Theory of Sampling applicable solutions have proven elusive in flowing pressurised particulate material streams, with various attempts to date not having formulated a viable, mechanically reliable and acceptable design.Compliance to sampling theory is critical, and a number of tests were conducted where certain operating parameters were altered to determine their influence and sensitivity on sampling representativeness.The results of the initial series of tests performed on a carefully designed rig are presented, with additional work discussed for necessary investigation in the near future.Getting closer to a robust, economic and correct design for this challenging application in the minerals industry may now not be that far away. Plant management staff will see improved metallurgical mineral balances around their process plants and significantly reduced sampling discrepancies, particularly in tailings applications. Improved quality control is now possible where, due to plant head room constraints, classical cross stream sampling methods cannot be employed.CITATION:Steinhaus, R C and Halstead, G J, 2014. True pipe sampler – a correct design utilising the principle of symmetry in pressurised slurry particulate streams, in Proceedings Sampling 2014 , pp 127–134 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 29, 2014

This paper cites selected, prominent examples of current, major projects in Southern Africa in both the mining and civil fields. Differences and similarities in the life cycle of mining and civil projects are discussed. A closer relationship between the mining and civil fraternities is beginning to develop and should be encouraged because of the resultant exchange of experience and technology which benefits both disciplines. An attempt is made to assess future potential for new, major projects and trends in the implementation thereof.

Jan 1, 1999

By McLennan T, Howells E, Zastawny H, Watt J. S

Nuclear techniques have been used to determine the solids weight fraction (W) and ash content of (fine) coal in slurries independent of voidage variations. The method is based on the correlation between W and the hydrogen concentration Hsl (wt/wt) of the slurry. Hsl is determined by combining measurements of neutron moderation and ?-ray transmission. Both Wand ash content are determined by combining the above measurements with the ash-sensitive measurement of X-ray backscatter compensated by a measurement of iron K X-ray intensity.The nuclear analysis system, called the SIRO coal slurry gauge, has been proved in a six-week plant trial at the BHP Steelworks washery at Newcastle, NSW. Several probes, each containing a radioisotope source and scintillation detector, were immersed directly into the coal slurry in the feed box upstream from the bank of flotation cells. For 140 samples W, ash and voidage varied in the ranges of 0 to 13 wt percent, 26 to 41 wt per cent and 0 to 4 vol. per cent, respectively. The r.m.s. difference in solids weight fraction, determined by probe measurements and conventional assay, was 0·52 wt per cent. Ash was determined with an r.m.s. difference of 2·0 wt percent in the coal. Better sampling techniques were used in the last half of the plant trial, and r.m.s. differences were 0.35 wt per cent solids and 1·5 wt per cent ash.

Jan 1, 1986

By T Baumgartl, D Mulligan, R Haymont

Mining at the Mt Leyshon Gold Mine in semi-arid north Queensland stopped in 2002. Newmont Australia has recently initiated a thorough post-closure water management study of the site by revisiting the existing information and conducting new water-related investigations. The focus of this paper, which is the first publication on post-closure environmental management of the site, is an overview of the site water quality in view of the sources and spatial distribution of polluted mine water, and also the performance of cover systems in controlling water flux though mine wastes. Since the mine closed the rehabilitation of waste rock dumps and tailings dams have significantly improved water quality in the surface receiving waters. However, the Scats Stockpile, the Heap Leach and Old Plant areas, which can potentially produce acidic waters, enriched with heavy metals and arsenic, are the top management priority. Modelling of the potential transport of contaminants in groundwater needs further investigations. A detailed examination of the waters in seepage sumps showed high concentrations of manganese, cobalt, nickel and selenium, which were previously overlooked. The study also showed distinctive chemical signatures for waters at different locations: the waters draining from scats, supergene materials and waste rocks have distinctly lower pH and higher concentrations of dissolved iron, sulfate, heavy metals, Mg, Mn, U, Se and As. While, the tailings sumps are more alkaline, and are enriched in Na and Cl. The southern tailing sump is characteristically As rich. The waters draining from the supergene materials have higher Cu concentrations. Assessment of the cover monitoring systems showed that although covers pass the design criteria, automated recording of data may ignore any failure of the monitoring system. Recalibration of the moisture sensors not only showed realistic moisture values compared to rainfall data but also showed that the permanently installed system records values that are higher than those of the mobile system. Event-driven recording of data overlooks infiltration events. Besides, with a higher time resolution (daily measurements) the breakthroughs could be clearly detected. Modelling of the water balance in covers is underway and an example is given here. Newmont is currently reviewing water management options and is willing to share the lessons learned.

Jan 1, 2006

By T Gready

Stawell Gold Mines operates the Magdala underground mine in western Victoria, Australia. Several orebodies are extracted by longhole open stoping at a depth of 900 m below the town of Stawell. Some stopes are backfilled with cemented slurry rock fill (CSRF) to ensure stability, increase recovery of ore reserves, minimise dilution of ore grades, and add value to the project. StawellÆs simple CSRF system is described, including the novel equipment used to minimise capital costs. Design criteria, operating procedures, monitoring systems and results are discussed in detail. Practical aspects are considered, including lessons learnt about barricades, overbreak and segregation. The economics of CSRF are discussed, including capital costs, operating costs and the total cost per ounce of gold poured. Options for mining with and without CSRF are compared in terms of return on investment and net cash flows. Examples demonstrate the method used to calculate æbreak-evenÆ stope ore grades which determine whether or not CSRF will add value in different circumstances.

Jan 1, 2002

By Mowat B. A

I lamata is within the Morobe Goldfield (7¦ 25' S, 146¦ 39' E) on the Wau (SB 55-14),1:250,000 scale map sheet (Fig 1). The deposit is 13 km SW of Wau, within the headwaters of the Upper Watut River, at an average elevation of 2000 m. The steep mountainous terrain is covered by dense tropical rainforest and access is by helicopter or by foot track. The deposit has been described previously by Wells and Young (1991) and 1)enwer. Leach and Mowat (1995). Data collected to late 1996 have been used to estimate an Inferred Resource of 9.2 Mt at 3.1 g/t gold using a cutoff of 0.5 g/t. This contains a high grade portion for which an Inferred Resource of 1.2 Mt at 9.7 grt gold has been estimated using a 5.0 g/t cutoff.

Jan 1, 1997

By P Jones

Fosterville Gold Mine is an underground operation located approximately 20 km northeast of Bendigo and 170 km north of Melbourne. Fosterville is the most productive gold mine in Victoria, producing over 100 000 ounces of gold annually. Access to the underground is via two portals cut into previously mined open pits. Fosterville is 100 per cent owned and operated by Northgate Minerals, a gold and copper producer with mining operations, development projects and exploration properties in Canada and Australia. (Northgate Minerals Corporation, 2010).Underground operations commenced in 2006 with ore production focused on three orebodies – Phoenix, Ellesmere and Falcon. Primary ventilation to the existing operation is provided by wall mounted exhaust fans located underground. This system provides approximately 220 m3/s of primary ventilation to the operations with intake achieved through the portals. Exploration drilling during recent times has uncovered another lode, the Harrier orebody, located 2.2 km to the south of the existing workings. Due to the location of the new Harrier orebody, vent analysis revealed the requirement for a separate source of primary ventilation to service Harrier. Figure 1 shows a current underground long section and the location of Harrier in relation to the existing workings. A number of options were considered, including a parallel decline from one of the existing two portals, however, a 4 m diameter, 350 m deep surface ventilation shaft was selected as the most suitable and feasible option for the mine.In December 2009, geotechnical drilling commenced to determine the most suitable location for the shaft. Site locations were constrained by a number of factors including interaction with current and planned surface infrastructure as well as proximity to nearby community members at the southern end of the mining lease. Results from the geotechnical drilling indicated that the majority of the proposed shaft would be suitable for a 4 m raise bored shaft. However, the top 40 m of the proposed shaft were deemed unsuitable for raiseboring due to zones of weakness that would likely influence the stability of a 4 m diameter shaft (Robinson, 2010). Following independent geotechnical advice and input from experienced raiseboring contractors, it was decided to construct a 40 m presink prior to raiseboring activities.A crane mounted auger was chosen as the preferred method to construct the presink. This method differs substantially from traditional blind sinking methods which are highly labour intensive. The auger, or drill, is designed and built to be attached to a crane base using engineered pivoting cradles. The auger is suspended from the crane jib by pennant ropes attached to hydraulic rams. The auger applies rotation and axial loads to the drilling tools via a telescopic steel bar (Kelly Bar). Drill tools are attached to the bottom of the Kelly Bar and comprise an assortment of drilling buckets and reaming barrels. Drilling buckets cut in situ material and collect the spoil for removal and are used for the excavation of the ‘pilot hole’. Reaming barrels are used to expand the diameter of the excavation by directing cuttings into the pilot hole which are then removed by the drilling bucket. In June 2010, the presink was constructed utilising a crane mounted auger with support to the shaft walls applied via a remote fibrecrete unit. The excavation method involved cutting a series of smaller diameter holes (2.4 m wide) before reaming out to the final diameter (4.7 m wide) in 4 - 5 m vertical increments. Fibrecreting activities were carried out on the backshifts, to ensure the stability of the excavation and minimise interaction between the two units. The presink was completed on time and offered many advantages compared to traditional blind sinking methods, including safety, costs and timing. This paper will examine the benefits of this method as well as lessons learnt and future applications in the mining industry.

Mar 21, 2011

By B P. Wrench, N R. Steward, C Wu

A gold mine in southern China has transitioned from an open pit operation to a fully underground mine. An overhand mining method is used which requires the mined stopes to be filled with backfill having an unconfined compressive strength of at least 0.35MPa. Tailings from the process plant is deposited into separate flotation and Carbon in Leach (CIL) tailings storage facilities. The tailings has a high fines content and the thickener on site could only produce a dilute flotation tailings at a solids content of about 30 wt%. This resulted in two undesirable outcomes: the density of flotation tailings in the Tailings Storage Facility (TSF) was significantly lower than planned so that the TSF could not store the life of mine tailings and that the backfill product had to include a high percentage of manufactured aggregate and increased cement content to achieve the required strength. This paper presents results of a tailings study which identified that the flotation TSF would be capable of storing the remaining life of mine tailings if an amended backfill product made up of 100% of flotation tailings could be introduced. A 100% tailings backfill with the addition of 8% cement was developed by classifying the tailings and using the coarser fraction in the backfill. This backfill allowed 60% of the flotation tailings to be placed underground and only the remaining 40% in the flotation TSF. A new thickener was designed and installed to thicken the tailings and rheological testing of the tailings, using a closed loop pilot plant, provided parameters to upgrade the underground reticulation system to deliver cemented backfill to stopes throughout the mine and to deliver thickened tailings to the TSF. The revised design was implemented and has resulted in considerable cost savings. CITATION:Steward, N R, Wrench, B P and Wu, C, 2018. Optimisation of surface tailings storage and tailings backfill – a case study, in Proceedings Mine Waste and Tailings Stewardship Conference 2018, pp 454–464 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 23, 2018

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 E Coffison, M Gibbs, C Bennett

The sustainable management of surface water associated with coal mining and coal seam gas projects is a considerable factor affecting the global competitiveness of these resource companies operating in Australia. Management of surface water can be costly, but effective water management and hence environmental stewardship enhances corporate social responsibility and the social license to operate. The surface water issues for the two industries are comparable, yet there is currently a lack of alignment between how Queensland, where the majority of AustraliaÆs coal seam gas projects are located, manages and regulates water on coal mine projects and coal seam gas projects.This paper provides a high level comparison of Queensland legislation and highlights the lack of alignment of this legislation regarding surface water management in these two industries. Through a case study, it looks at the surface water management process at a coal mine site based on the current legislation and principals of surface water management. QueenslandÆs lack of alignment in legislation regarding surface water management in these industries is understandable given the different development paths and time scales of the two industries. The coal mining industry has developed over a long period of time and the legislation has progressed in parallel, whereas, the coal seam gas industry is relatively new in Australia and the legislation has had to advance quickly. The lack of alignment between management of surface water for the two industries creates confusion among stakeholders and this adds complexity and additional transaction costs during the exploration, permitting and operational phases of projects, including landowners who may be affected.CITATION:Bennett, C, Coffison, E and Gibbs, M, 2013. Lack of alignment between regulation of surface water management in coal mining and coal seam gas, in Proceedings Water in Mining 2013 , pp 157-162 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 26, 2013

By J-P Franzidis, E V. Manlap

Froth transportation characteristics have a significant impact on the flotation recovery especially in large flotation cells. The probability of an attached particle being recovered into the concentrate depends strongly on the location where it enters the froth phase. A mathematical model for froth transportation in industrial flotation cells was proposed (Zheng et al, 2004). The model has been tested on the experimental data obtained from testwork conducted at the Western Metals Hellyer lead/zinc concentrator. The results show that the model fits the experimental data well, especially when the froth is well developed (ie for a deep froth). The model fitting can be significantly improved by incorporating an air dispersion factor. However, the models encounter some limitation, ie inconsistent prediction at different froth heights, due to change in froth viscosity which cannot be modelled at the present time.

Jan 1, 2005

By A Moreira, A B. De Souza, S Latchireddi, E Faria

Material transport plays a significant role in the efficiency of horizontal grinding mills - autogenous (AG)/semi-autogenous (SAG)/ball mills. The magnitude of benefits that can be achieved with efficient material transport in AG/SAG mills has been demonstrated in a greenfield milling operation at Mirabela MineratpoÆs SantaRita Nickel operation located in north eastern Bahia in Brazil. This grinding circuit was commissioned in October 2009, was designed as a typical SABC circuit to grind 4.6 Mt/a (575 t/h) to produce a P80 of 125 ¦m with ten to 12 per cent ball charge in the SAG mill. However, it has been operating as an ABC circuit without any grinding media in SAG mill. The nameplate design capacity was achieved in few days and was gradually increased to an average of +800 t/h, still operating in ABC mode. The Santa Rita grinding circuit consists of one 30 ft + 15.4 ft 8.5 MW SAG mill fitted with Outotec patented Turbo Pulp Lifters (TPLÖ) in closed circuit with a HP400 short head pebble crusher and one 20 ft + 28.4 ft 5.5 MW Ball Mill in closed circuit with a Warman cyclone cluster fitted with 10 + 26 in CAVEX cyclones. This paper examines the comminution design process, operating philosophies, experiences, post commissioning optimisation and future development works. Evaluation of the milling circuit performance relative to the design criteria is also discussed.CITATION:De Souza, A B, Latchireddi, S, Moreira, A and Faria, E, 2012. Application of turbo pulp lifter in greenfield operation - Mirabela SantaRita nickel operation, in Proceedings 11th AusIMM Mill OperatorsÆ Conference, pp 53-58 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Oct 29, 2012

By R Crosbie, K C. Runge, E Tabosa, J K. McMaster

Flotation feed density is an operating parameter which is currently not well understood. Most models of flotation only incorporate its effect on particle residence time, but it is likely to have additional effects within both the pulp and froth phases of the machine. Test work was performed in MetsoÆs 3 m3 RCS test rig to investigate the effect of flotation feed density on metallurgical performance. The rig was operated at Rio TintoÆs Northparkes operation using flotation feed and scavenger tail. For each feed, tests were performed at different feed densities, and measurements were performed to estimate the pulp and froth recovery as well as other key flotation parameters (eg bubble size, turbulence and froth stability). Results of this test work have shown that the pulp density does affect key parameters of the flotation process in both the pulp and froth phases. Pulp phase kinetic rates decrease at high density and this is attributed to a reduction in the turbulence within the cell. Froth phase recovery, however, improves presumably because of a stabilising effect due to a greater loading of solids on bubble surfaces.There is a need for more research to be done to better understand these interactions so that metallurgists can determine the optimum flotation feed density to use at a particular operation, or within a section of a flotation circuit.CITATION:Runge, K C, Tabosa, E, Crosbie, R and McMaster, J K, 2012. Effect of flotation feed density on the operation of a flotation cell, in Proceedings 11th AusIMM Mill OperatorsÆ Conference, pp 171-178 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Oct 29, 2012

By R Osterman

As mining companies have to go deeper in search of ore and coal, the importance of protecting the highwall increases. Wall fatigue also increases with importance of the long life expectancy of a highwall, in particular the impact of fatigue influenced by resonance. Resonance is the tendency of a material to oscillate with greater amplitude at certain frequencies when exposed to a periodic force. Until recently, only dominant frequencies, either natural or induced, were all that could be measured through the fast Fourier transform of the blast-induced wave form.This paper presents a novel method to accurately predict the resonant frequency of a highwall. In fact, the method can predict the natural frequencies between any two monitoring points placed in line with the source of impact, and as such it can be used to determine the resonant frequencies of any physical object. The method not only predicts the resonant frequencies of the material; it predicts the structural response of the total frequency spectrum. This method helps describe the response of a mechanical structure to excitation. It describes the structure’s efficiency in the use of energy, that is, whether energy is dampened or consumed in resonance.The initial predictions have been so accurate that it made it possible to determine the fundamental frequency and the first six harmonics of a highwall in a northern Queensland coalmine. With the use of electronic detonators, the blast can be timed with a repeating delay to create an induced artificial frequency that has been determined to retain the least amount of energy.CITATION:Osterman, R, 2015. Transfer function – new method to accurately predict resonant frequencies of a highwall, in Proceedings 11th International Symposium on Rock Fragmentation by Blasting, pp 233–244 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 24, 2015

By Beamish B. B

In this study, a systematic examination of the scale and variety of micro-structures in coal using Scanning Electron Microscopy (SEM) has been carried out, using coal samples of differing stages of coalification from the Bowen Basin and New Zealand. SEM provides detail of various micro-structural features and shows that a variety of micron-sized fractures and cavities are present in coals. These are differentiated into three pore types: fracture porosity, phyteral porosity and matrix porosity. Fracture porosity is generally associated with bright coals (although it has been shown that micro-fractures are present in maceral fragments from the dull coal layers). Characteristically the macro- and micro- fractures form a continuous structural fabric through the bright coal layers and thus make a significant contribution to overall permeability. The macro- and micro-fractures appear to have been formed during the process of coalification rather than being formed due to shrinkage during methane extraction, as most of them are infilled with secondary minerals. In contrast, phyteral and matrix porosity is associated with the dull coal layers that are composed of plant fragments or a heterogeneous mixture of macerals. Like the macro- and micro-fractures, the continuity of the observed cavities suggest that these too are a significant contribution to overall permeability, and therefore play a major role in the transmissibility of methane at a level between diffusion at the micropore level and laminar flow at the macro-fracture level. Due to different spacing, orientation and persistence of individual fractures and cavities, the rate of methane transmissibility through a coalseam will vary according to coal rank, coal type and the organic precursors from which the coal was derived. This is important as a knowledge of the micro-structure system and its relationship to coal type and coal rank, plays an important consideration in: a) gas drainage modelling, b) enhancing permeability by inducing fractures and c) predicting the occurrence of outbursts.

Jan 1, 1991