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By N Eriksson

BolidenÆs Aitik mine, located 100 km north of the Arctic Circle in northern Sweden, is one of EuropeÆs largest copper mines. This operation produces 18 Mt of ore per year and 99 per cent of this tonnage needs to be deposited as tailings. The tailings pond currently covers 1140 ha but at closure it will hold around 760 Mt of tailings and occupy an area of 1400 ha. The tailings contain less than one per cent sulfur and only 0.03 per cent copper. The carbonate content is as low as 0.25 per cent but silicates, with substantial accumulated buffering and metal adsorption capacity, are abundant. A programme of comprehensive material characterisation and investigations has been undertaken in order to design a decommissioning strategy that meets the regulatory requirements and secures long-term physical and chemical stability of the deposited tailings. The main concern has been the potential generation of acid rock drainage (ARD). Standard tests using acid base accounting (ABA) and humidity cell tests during 15 weeks were not conclusive regarding the ARD generation potential. The accuracy of the results was considered to be unsatisfactory, therefore the humidity cell tests were gradually extended in time. In total, almost four years of humidity cell tests were conducted on various sets of samples which were selected in order to cover the variation in sulfide content existing on site and in various points of the milling process. The laboratory results were subject to detailed geochemical evaluation in order to fully understand the interaction between acid-producing reactions, driven by weathering of sulfides, and acid-consuming reactions by dissolution of carbonates and alumino-silicates. The laboratory tests were followed by both pilot tests and field measurements in order to estimate the scale dependency between the laboratory and the field results. It was found that sulfide weathering rates in the field were considerably lower than those measured in the laboratory. However, it could not be verified that the tailings top layer would maintain a neutral pH in the long term even under field conditions. Thus, a reviewed tailings management plan, including de-pyritisation of the tailings during the last years of operation, is considered.

Jan 1, 2003

By S Bourke, R Stone, B Hill

Since Jameson Cell flotation technology was first installed in a coal flotation application in 1988/89 at Newlands Coal Handling and Preparation Plant (CHPP) there have been fundamental improvements to its design which has led to its current status as the leading technology in Australia for fine coal flotation. These improvements, which include key hardware modifications, consist of the re-development of the orifice plate to produce considerable increase in wear life and reduction in power consumption for a given air entrainment and a significant increase in the amount of air that is entrained. Modifications to the downcomer allow maximisation of residence time and air entrainment while changes to the feed and air distribution system allows a reduction in installation cost and a reduction of solids ingress into the air line. Additionally, operational improvements such as the use of recycle to maintain constant flow to the cell feed are discussed. The recent installation at Hail Creek in the Bowen Basin, which consists of three B6000/20 Jameson Cells (6 m diameter bottom-fed distributor with 20 downcomers) will be reviewed as a case study. The latest design and operability of the cells are summarised and compared against one of the original designs at Riverside CHPP also in the Bowen Basin.

Jan 1, 2005

By G E. Smith, G N. Lye

Wall control blasting practices are necessary to reduce the impact of blasting on mine faces but can also have a significant negative impact on mine productivity and operating costs. The conventional practice in deep open pit mines is to use so-called trim blasts adjacent to pit walls. To provide burden relief these trim blasts have fewer rows than full production blasts and are fired to a cleared free-face; hence they are termed æunchokedÆ. This practice leads to scheduling constraints on the pit operations and can cause ore dilution due to excessive muckpile movement. The use of such trim blasts stems from the perception that increased wall damage results from æchokedÆ blasts. These concerns are based on the unproven assumptions that blast vibration levels and explosive gas penetration increase with increased blast burden and face confinement. This paper describes work undertaken as part of a major investigation into wall control blasting at the KCGM Fimiston Mine, Kalgoorlie, Western Australia. It details a study to assess damage effects due to blast burden. Borehole air pressure measurements and borehole video camera inspections were done behind a series of single blastholes drilled with varying burden distances, as well as behind a dedicated trim blast and a full production blast. It was found that the measured damage effects, including visible rock cracking, dilation, and the limited extent of gas penetration behind the blastholes, did not vary significantly with burden or blast type for the cases tested. This result was in complete agreement with detailed vibration measurements conducted by Blair and Armstrong (2001) during the study, which found that vibration was independent of blast burden. As a result of these investigations, changes to the blasting practices at the mine were implemented. Dedicated trim blasts and free-face blasting have been replaced by modified production blasts and the practice of æchokingÆ blasts has been introduced. This has resulted in a significant improvement in productivity and cost savings without compromising pit wall integrity.

Jan 1, 2001

By W P. Darcey

Jabiru Metals completed a feasibility study on the Jaguar Base Metal Project in July 2005. The study was based on conventional truck haulage via a decline. Conveyor haulage was considered but not in sufficient detail to provide a meaningful comparison. The current study investigates the cost and productivity of alternative ore transport systems for the Jaguar project. The options considered are conventional underground truck, conveyor, road train, rigid frame truck, monorail haulage and shaft hoisting. Shaft hoisting was not considered in detail due to the high capital outlay for such a relatively short mine life of less than six years. This study goes beyond the common approach of evaluating haulage systems in isolation; it includes development costs, timing of ore production and costs associated with mine services as well as administration and milling costs. The results of this study indicate that the best haulage option for the Jaguar project is conventional truck using the Hitachi AH400 trucks. The net present value (NPV) for this option is A$8.13 million higher than the next best option namely the Kiruna E635 electric truck. The conveyor option has the lowest operating cost per tonne of ore but the longer development time negatively impacts on its NPV. The monorail option is considered impractical due to the number and length of trains required to achieve production targets.

Mar 21, 2011

By Warren L. J

The recovery of phosphate minerals from the Mt Weld deposit by froth flotation has been studied by many researchers. Most of the flotation experiments were carried out with a tall oil fatty acid as collector and sodium silicate as dispersant and gangue depressant at pH about 10. Several rougher stages and cleaner stages and, sometimes, a desliming stage were included. The major difficulty was to reduce the Fe2O3 impurities in the concentrate to the required level. We used QEM*SEM (Quantitative Evaluation of Materials by Scanning Electron Microscopy) to provide a comprehensive and accurate mineralogical analysis of the flotation products. The analysis indicated that the Fe2O3 impurities were mainly recovered in the second concentrate, owing partly to the incomplete depression of the liberated iron oxides by the sodium silicate and partly to composite particles containing iron oxides. The selective depressing action of sodium silicate was improved by adjusting the variables such as the modulus (SiO2/Na2O), pH and time of standing of the stock sodium silicate solutions or by adding some polyvalent metal salts into sodium silicate solution. The flotation process was optimised to produce a phosphate concentrate with the required grade and recovery (38.6 per cent P2O5, 2.67 per cent Fe2O3, 13.7 for P2O5/(Fe2O3 + A1203), at P2O5 recovery of 81.5 per cent). The possibility of recovering the rare-earth phosphate, monazite, from the Mt Weld deposit by flotation was also studied.

Jan 1, 1991

By D Kahler

Modern gold production from the Gympie Goldfield commenced in 1995 with the first production from the Monkland Mine. The primary ore target was the 2.5 m wide Inglewood Fault ore zone, which was initially mined using traditional narrow-vein shrink stoping. This manuallyintensive mining method allowed the mine to re-open with minimal capital cost, however the trade off was low productivity and high operating costs. As the mine has developed, access through 100-year-old, timber-lined, small shafts, and the high cost of delineating reserves, has restricted modernisation of the operations. In early-2000 a stockwork lode was discovered adjacent to the existing workings with an economic width of 40 m. In order to maximise the benefit of such an orebody, mining needed to be approached in a new way. This paper summarises the many changes implemented at the Monkland Mine, including the necessary cultural change, implementing long-hole stoping, de-bottlenecking and the application of mechanisation. The result has been a significant improvement in safety performance, higher productivity and a more secure future for all those involved in the project.

Jan 1, 2002

By A H. Gray

Gekko Systems Pty Ltd engaged in the test work, design, manufacturing and commissioning of the pre-concentration plant at Silver Standard Resources Inc’s Pirquitas mine in the north western Jujuy province of Argentina. Testing of the ore at coarse feeds of between 2 and 12mm returned probable recoveries of up to a total of 95% silver into 50% of the mass fed to the InLine Pressure Jigs (IPJ’s). The plant has been designed and built based on the test results to include a preparation screen to bypass the fines that are naturally higher in grade and repulping of the coarse which is then pumped to three parallel trains of 2-stage IPJ roughing-scavenging. The target 50% yield to concentrate passes over a dewatering screen and the solids are returned back to the mill feed conveyor. The IPJ tails are transferred to a dewatering screen and solid rejects are stockpiled off the end of a transfer conveyor. All dewatered underflow products are processed via hydrocyclones to recover dirty water for use back into the IPJ circuit and return entrained fines to the fines bypass line.With the sulphide orebody being processed, the results from operations to date have shown that plant performance is matching the original laboratory sulphide ore testwork data. The objective of the pre-concentration step is to both reject gangue and achieve grades between 300-400g/t Ag to the float circuit as flotation recoveries are maximised within this range. The plant currently achieves overall upgrades of approximately 150% of feed grade with the IPJ only circuit producing an average 180% increase from calculated feed grade.

Aug 8, 2011

By D L. Reid, G Nwaila, J Petersen, J-P Franzidis, Y Ghorbani, M Becker, F C. de Beer, L C. Bam

Traditional methods of gold extraction have involved comminution and processing via cyanide leaching, all of which are associated with energy, cost and environmental implications. Subsequently, numerous opportunities for improvement and optimisation strategies in gold mining and processing have been explored, including heap leaching, which has significantly lowered comminution costs. This paper explores the effect of the more energy efficient high-pressure grinding rolls (HPGR) on the comminution behaviour (especially on the crack density) and cyanide leaching performance of laboratory reactors study on run-of-mine gold ore comprising the Ventersdorp Contact Reef (VCR) and the Elsburg Reefs (ER), sourced from Carletonville in the Witwatersrand Basin in South Africa. The ore was comminuted using HPGR at three different pressure settings (120, 90 and 60 bar). Large (-16/+11.2 mm) and small (-5.6/+4 mm) particles of the HPGR product where then mineralogically characterised using optical microscopy, quantitative evaluation of minerals by scanning electron microscopy (QEMSCAN) and X-ray computer tomography (CT) before cyanidation. The results of the leach experiments indicated a correlation between gold extraction, particle size and crack density. Mineralogical characterisation of the ore indicates that the Fe-sulfides present in the VCR are likely to be more reactive than those in the ER, thus having a more favourable influence on gold cyanidation. Overall, the results show promise for the heap leaching of Witwatersrand gold ores prepared using HPGR and demonstrate the value of a multi-disciplinary geometallurgical approach.CITATION:Nwaila, G, Becker, M, Ghorbani, Y, Petersen, J, Reid, D L, Bam, L C, de Beer, F C and Franzidis, J-P, 2013. Morro do Ouro geological model with a metallurgical view, in Proceedings The Second AusIMM International Geometallurgy Conference (GeoMet) 2013 , pp 75-84 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 29, 2013

By J Janus, J Krawczyk

Local accumulation of methane in workings of a longwall is one of the most common causes of mining accidents. This gas has many sources: it may leak from goaf through permeable sealings, or through gaps between roof support sections. This stream is then mixed with ventilation air. In the case of auxiliary ventilation ducts, or free jet fans, we have to consider the propagation of a clean air stream in areas with an air-methane mixture. Therefore, modelling the mixing process of methane with ventilation air is important for determining the hazardous areas and thus to ensure safety. Computational fluid dynamics (CFD), specifically the finite volume method, provides tools for such modelling. However, this method requires validation. For this purpose, we have compared measurements with simulations for selected test cases having sufficient similarity to the present problems. In typical methane explosion hazard areas, such as a longwall, it is difficult to make measurements sufficiently to confirm credibility of 3D CFD simulations. As a relevant alternative, an injector methane drainage station placed in an underground gallery has been selected. The stream of drained goaf methane is entrained by compressed air powered injectors and then mixed with ventilation air. Research in this area does not interfere with mining operations, it can be performed safely in a repeatable way. Flow within the site has many similarities to conditions in mine drifts of the longwall area. Comparative research of measuring the methane concentration distribution in the mine drift downstream injector station, with simulations of propagation of the methane stream, was carried out. For measuring the methane concentration distribution in the mine drift a multipoint measurement velocity system developed at the Strata Mechanics Research Institute was used. For the simulation a finite volume method for turbulent flow of gas mixtures was applied. Preliminary results of comparison are shown as graphs and tables.CITATION:Janus, J and Krawczyk, J, 2013. An analysis of the mixing of air and methane in the stream produced by the mine injector station - Presenting preliminary results of measurements and modelling, in Proceedings The Australian Mine Ventilation Conference , pp 49-56 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 1, 2013

By M Thompson

"The Jaguar concentrator, 60 km north of Leonora, Western Australia (WA), consists of a standard closed circuit semi-autogenous grinding (SAG) and ball mill producing a float feed of approximately 60 micron at 70 t/h. Following comminution, separate copper and zinc concentrates are produced by differential flotation. These concentrates are subsequently dewatered by thickening and then filtration.Separate audits of the Jaguar Operation’s copper and zinc concentrate thickeners were conducted to evaluate their performance, following occasions where site personnel had identified unnecessary loss of copper and zinc concentrate to the thickener overflow. The feedwells were identified as the source of the observed problems due to the following issues with their design/duty:lack of retention time in the feedwell as current design is completely opennot conducive to effective mixing required for efficient flocculationdiameter of the zinc concentrate thickener feedwell was undersize for its duty.To improve the performance of the thickeners, it is planned to replace both concentrate thickener feedwells with a high rate design for their respective duties in February 2016. The audits also identified that the feed system requires modifications with particular importance placed on minimising air entrainment of slurry into the thickener feedwell. Minimisation of air entrainment is hoped to be achieved by providing constant head pressure through installation of weir arrangements in the feedboxes and redirecting each feed pipe so that it feeds tangentially into the feedwell. In addition, improvements to the control philosophy of both thickeners will be made with the installation of an ultrasonic bed level probe and bed mass sensor, in order to make real-time adjustments to the underflow pump speed to maintain steady state operation.Following the feedwell replacement and modifications to both thickeners it is expected that losses of recovered copper and zinc will decrease to undetectable levels. Subsequently it is anticipated that a more consistent, homogenous underflow can be produced for downstream filtration.Citation:Thompson, M, 2016. Concentrate Thickeners Feedwell Replacement, in Proceedings 13th AusIMM Mill Operators’ Conference 2016, pp 273–278 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Oct 10, 2016

By H Asche

The Melbourne Metro project has evolved from the Victorian Transport Plan. MM1 is the second major phase of the proposed separation of Melbourne’s radial railway system, creating a line that will operate independently across the metropolitan region. The central section will run beneath the inner suburbs and the central city. The initial project comprises nearly 8 km of twin bore tunnels, three cavern stations and two cut and cover stations, as well as a dive structure and connections with the existing rail system. This paper describes the development of the geological model, a modification of the rock mass classification system derived during the EastLink project and the influence of the geology on the design of the tunnels and stations for the Department of Transport’s business case.The majority of the tunnelling will be in the siltstone and sandstone bedrock. This material has been extensively folded, sheared and intersected by igneous intrusions. A revised rock mass classification system has been developed for the Melbourne Formation to assist in identifying the important engineering properties of the rock and to rationalise the support types required.The Metro’s tunnel alignment crosses the Yarra Delta deposits in the lower reaches of the Moonee Ponds Creek and again at the Yarra itself near Princes Bridge. These deposits vary in extent and thickness over very short distances. The investigation of these materials has been carried out using conventional drilling techniques in conjunction with sonic drilling methods.Over the central section of the Metro, the positions of the new station are determined largely by transport planning requirements and urban development opportunities. However, the stations, and the tunnels, have been the subject of studies, optimising exercises, and selection processes to arrive at the concept design. This paper describes the influence of the varied geology on the development of the design, and the identification and management of the risks in tunnelling under the centre of Melbourne.

Mar 8, 2011

By F Soleimani

Coal and gas outburst is a common hazard in coalmines and due to the complexity of its nature, efficaciously predicting the possibility of its occurrence is intricate. The mechanical and gas flow properties of coal, in situ and mining-induced stresses and geological structure are involved in outburst occurrence, whereas suggested thresholds for its prediction do not consider them all. Numerical modelling is known as an effective method that can not only consider couple the effect of different physics but also analyse the effect of associated parameters. In this paper, a coupled fluid flow and geomechanics model has been developed in COMSOL Multiphysics to simulate outbursts by considering the effect of gas pressure and composition, damage mechanics on related coal properties, and progressive excavation and mining-induced stress. For the proposed fully coupled outburst model, the effect of pore pressure and sorption are applied to the stress while the influence of stress changes and damage are considered in updating coal porosity and permeability in each time step. In addition, to study the effect of mining-induced stress, the excavation process is simulated by changing the material properties over time according to the schedule of heading development. Then, outburst initiation is analysed from the energy point of view by using the concept of energy conservation law. The simulation results are used in calculating energy sources of the outburst initiation, also, the effect of particle size and ejection velocity are studied, and the significance of particle size effect on the energy analysis and deformed coal on outburst initiation is discussed.

Oct 12, 2022

By M Sebbag, T Taimre

Barrick Gold Corporation is proceeding with development of the Buzwagi Mine in the Lake Victoria Greenstone Belt District of Tanzania. The Buzwagi orebody occurs in the Nzega Shear zone within the Greenstone Belt and is hosted in a K-Feldspar rich granite in contact with mafic volcanic. The orebody presents 2.6 M ounces in 41 Mt of reserves, at 2.0 g/t. Over the ten year mine life, total material movement will be 261 Mt. The feasibility study was completed in 2006 and presented several technical challenges. Two principle challenges are grade control and the impact of dilution on the economic viability of the project. The principle challenges for grade control will be quality control and quality assurance in the sampling of the massive coarse grained granite, which comprises 70 per cent of the deposit. The principle challenge with dilution will be with the selective mining of the narrower mineralised zones contained within the mafic volcanics. This paper will focus on the mine planning methodology which has been applied including selective mine unit (SMU) and dilution estimation, net smelter revenue (NSR) calculation and the effects of cut-off grades to the project economics and mine schedule.

Jan 1, 2007

By R Bourcier, T Rafferty

Following an underground coal mine explosion, it is important that ventilation is quickly restored to ventilate the mine and affect a mine rescue. To achieve this, surface ventilation fans must be suitably protected from explosion damage, but still able to be quickly restored.A review of all Australian coal mine regulations was carried out, revealing that Australia has no prescriptive regulations that specify the design of explosion relief panels and the protection of ventilation fans in the event of an underground explosion.To better define the level of protection required, this paper reviews and extracts details from relevant major international regulations and standards on the subject.A suitable design methodology for the protection of ventilation fans is presented that could be applied in Australia, using a combination of explosion relief panels, appropriate isolation of the fan, and ventilation shaft/mine design.CITATION:Rafferty, T and Bourcier, R, 2013. The protection of surface ventilation fan installations from underground explosions, in Proceedings The Australian Mine Ventilation Conference , pp 169-173 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 1, 2013

By S Liu, N M. Rana, A Take, N Ghahramani, D Clohan, S McDougall, S G. Evans, H J. Chen

Tailings dam breaches have caused significant life loss, social, economic and environmental damage in recent years. The term "tailings-flow" refers to various forms of tailings outflow movement resulting from the breach of a tailings dam. These flows, comprising a mixture of tailings, water and dam materials, can take different forms, ranging from a debris flood to a flowslide, and can travel long distances and impact large areas. The inundation extents depend on the failure mechanism, tailings dam height, total released volume, sediment concentration (associated with the rheological behaviour of the flow) and downstream conditions (including topography, land cover and presence of a water body). Numerical runout modelling of tailings dam breaches is important for predicting the dynamics and extent of tailings-flows, which are crucial steps in tailings dam consequence analysis and risk management. This paper presents a comparison of numerical simulations of the 1994 Merriespruit tailings dam breach using three numerical models: DAN3D, FLO-2D and MADflow3D. Using compiled site condition information and data from the literature and public domain, model performance was evaluated primarily based on comparisons between the observed and simulated inundation extents. The main purpose of this preliminary model benchmarking exercise was to help us evaluate the capabilities and limitations of each model, and to identify opportunities for future model improvements as part of a broader research initiative known as CanBreach. This exercise also highlights the inherent complexity of tailings-flows and the importance of using judgement to interpret model input and results.

Jul 1, 2021

By H J. J Boshoff

Mine closure planning integrated into life-of-mine (LOM) planning during the operational phase consists of many subcategories. One of these subcategories that needs to be fully integrated into the LOM planning is that of water management planning. Issues arise when mine water management infrastructure, which is required post-mine closure, is only considered closer to the end of LOM. Once the mine is closed monitoring is focused on water quality as an indicator of closure performance, hence inadequate upfront post-mining water management planning or integration into operations will result in environmental harm. The lack of considering post-mining water management is not uncommon in LOM planning. Poor mine site water management and planning are a wide reaching problem for the mining industry. Numerous active mines have been issued with environmental notices and penalties for contaminated water discharge; including issues with abandoned mines now being cared for by governments all over the world.Mitigating sources of water contamination caused by existing and closed mine sites that are leaching pollutants might be impractical, due to inaccessibility to the contaminants or final landform topography. The opportunity to consider and introduce post-closure required mine water management infrastructure is lost once the waste facilities are constructed. Mitigating contamination is made more problematic with mines located in remote areas, serviced by dedicated roads, water supply, electrical supply, etc. It is expected that these services will be decommissioned or drastically reduced, once closure is reached making remedial work much harder. Inadequate planning, to ensure access and easily maintainable water management infrastructure, could lead to erosion and contamination going unnoticed. There are numerous examples of poor mine water management practices in Africa, Asia, Australia, South and North America. In the United States more than 23 000 km of streams are affected by pollution sourced from some 50 000 abandoned or mature mine sites (DeGraff, 2007).CITATION:Boshoff, H J J, ?2014. Inadequate water management planning resulting in long-term closure liabilities, in Proceedings Life-of-Mine 2014 , pp 233–246 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 16, 2014

By P Viljoen, A Holland, C Parken

In the raisebore industry, there are two known methods of removing large diameter reamer heads once the reaming of an underground shaft is complete: 1. Lower the reamer head to the base of the shaft and use a thermal lance to cut the drill string and drag the reamer away from the hole. This method exposes the person doing the job to falling rocks from the unsupported mineshaft, a hazard which is difficult to control and has caused fatalities. 2. Install breast plates in the backs above the reamer head at the top of the hole, and use air hoists to lift the reamer out of the hole. Steel beams are then placed across the hole which the reamer is lowered down onto, dragged forward away from the hole, disassembled, and removed. This method is time consuming, expensive, and complex in the controls required to ensure safe completion of the job At Carrapateena, the air hoist method had been successfully used, however, with the cost involved and the complexity of this process, development of an improved and fit for purpose reamer removal method was pursued. OZ Minerals collaborated with Pybar to develop an industry first application of Enerpac SL100 lift and shift technology; a track mounted system with hydraulic lifting units that can both lift the reamer out of the hole and tram it back along the tracks away from the hole. A key component of turning the idea into a reality was by establishing a truly collaborative relationship between client and contractor. When the SL100 mobile gantry arrived, it was set up in Adelaide and trialled with the whole raise bore crew attending the two-day workshop and risk assessment; a rare opportunity for drillers. There have now been five successful reamer removals with the SL100 mobile gantry at Carrapateena. It has proven to be an efficient process with the entire process from setup to demobilisation taking less than 48 hours. This new way has completely removed hot works and uncontrolled hazards from working at the bottom of a shaft. It has also saved the reamer stabiliser from being cut and thus wasted. Exposure to the open hole has been minimised significantly as assembly of the SL100 is completed well forward of the open hole. This new process is exponentially safer, quicker and cheaper.

Mar 15, 2021

By A Van Kersen, J Hodson

Waihi Gold Mining Co Ltd [WGM], operated by Normandy Mining, has been internationally recognised for its achievements in environmental management at the Martha Mine at Waihi. In a proactive approach to dealing with community concerns regarding blast vibrations on the North Wall cutback, trials were performed using the Cardox System [Cardox], to determine feasibility of achieving economic fragmentation of andesite overburden whilst bettering management levels of 5 mm/s peak particle velocity. Cardox is a non-explosive pressure breaking system which uses high pressure CO2, Results against conventional Ammonium Nitrate/Fuel Oil [ANFO] using a variety of delays showed that conventional blasting techniques could produce marginal fragmentation within the management levels with occasional excursions, whereas Cardox produced the required fragmentation while not registering on the vibration monitoring equipment. Fragmentation to enable normal excavation by a Hitachi EXII00 hydraulic excavator was achieved with a yield of 16m3 per round. The trails indicated that fragmentation with Cardox at the yields required of over 18m3 per round was possible with further product development. The trial demonstrated that non-explosive pressure breaking is feasible and its application warranted where environmental concerns affect the cost-benefit economics of production blasting in sensitive areas.

Jan 1, 2001

By Bidwell TJ, Servant S

Paste back-fill technology is gaining rapid acceptance in the Canadian mining industry. The authors were able to work on four different mining sites employing paste technology in their operation. In each site the paste hackfll is prepared with very fine tailings containing significant amounts of sulphur minerals. The research program involved laboratory testing on samples prepared in the laboratory, on control samples obtained at the hackfilll plant during pouring into the stopes and on core samples obtained from drilling in backfilled stopes. The laboratory testing provided detailed knowledge of the mechanical behavior of the paste backfill. The need for a means of measuring the in-situ mechanical properties of the backfill became clear. A review of various trials published in literature and an evaluation of available technology suggested the Cambridge Self-boring Pressuremeter. A first field campaign confirmed the efficiency and dependability of this instrument leading to the conclusion that it offers a very good method for the in-situ measurement of paste backfill properties.

Jan 1, 1998

By N Winkelmann

Underground coal heating is a significant hazard in some coal mines. The broad principles of how underground coal heatings occur is reasonably known, however the technique for locating the heatings in often inaccessible goaf has been proved to be difficult. Since 2002, CSIRO, with the support of the Australian Coal Association Research Program (ACARP) has been undertaking investigations to apply and develop innovative surface-based radon technique for locating the underground heatings in Australia. This paper presents the progress of the investigation.

Jan 1, 2005