Search Documents

By I Sucholutsky, D M. First, D Mogilny, F Yusufali

Machine learning is creating value in all facets of the mining industry, from exploration to production. The authors provide an accessible, high-level introduction to artificial intelligence (AI), machine learning (ML), and deep learning (DL); the latter being widely recognised as one of the most powerful forms of ML. In particular, the authors will introduce deep learning models known as convolutional neural networks (CNNs), how they are applied, and the economic considerations necessary for determining when DL may be the right solution to de-risking complex block modelling problems. The authors present preliminary results from the mineral resource modelling study of the Jundee orogenic gold deposit, Yandal belt/gold province, Western Australia. The primary goal was to identify the direction and location of the narrow gold veins more accurately and demonstrate how non-linearly correlated elements are used as direct inputs to the resource model to assist with the target element’s grade prediction. This demonstrates that: (1) existing techniques for finding correlations between assayed elements do not adequately reflect the complex geology of the asset, (2) nonlinear correlations that are difficult to model as simple mathematical functions are representative of geological patterns in a deposit, and (3) non-linearly correlated assayed data, fed as inputs, increase the performance of the resource model as reconciled through blind tests. To conclude, the authors hypothesise that the patterns represented by the DL block models may be revealing the results of overprinting geological processes that generated mineral deposits. For example, the primary hydrothermal processes that deposited metals created depositional patterns that become particularly complex as a result of being overprinted, in part or in whole, by secondary physicochemical processes. This may explain why non-linear geochemical relationships have the capacity to generate more accurate block models of mineral deposits.

May 24, 2023

By M Gill, A Mathias

The mining industry has always been cost-focussed but in recent times mining companies have moved towards obtaining a competitive advantage through the implementation and sustainability of best practice in the areas of improved safety, operating costs, quality and productivity. This is the concept of Value In Use û that mining companies see value generated all the way to the bottom line through improved change, and those changes become an established mine practice. Many mining companies do not have the resources or the knowledge to sustainably achieve such an outcome. This has led to companies looking to their suppliers to deliver measured value to their business through partnership arrangements. Suppliers have been able to leverage their knowledge of products, product applications and services to offer alternative supply agreements that bring greater returns to their businesses while improving supplier/client relationships. Orica has developed such a partnership with Coal & Allied in the form of a Continuous Improvement Program. Coal & Allied is based in the Hunter Valley region of NSW and is the largest coal producer in that region. Both parties are contractually bound to jointly deliver Value In Use projects to Coal & AlliedÆs Hunter Valley Operations and Bengalla mine sites. It is the request of the Orica and Coal & Allied Steering Committees that this paper does not include the details of actual projects or the costs or key performance indicators associated with the program. This is to maintain the competitive advantage to both Orica and Coal & Allied through the implementation of such a program. It is in the writersÆ opinion that this paper does not lose any significance through this restriction as the most important component of the Continuous Improvement Program is not the projects themselves but the methodology behind them. With this in mind the paper will discuss the development of Value In Use offerings, the key aspects of the Coal & Allied and Orica joint program, what is a Value In Use project, how to measure its value, finally why and even more importantly what makes this program is a success.

Jan 1, 2004

By Vestre J, Kure K, Bergqvist I

Existing practical laboratory tests for comparative strength determination of commercial explosives consist of the lead block test, the ballistic mortar test, the bubble test etc. These tests are carried out in material with quite different properties compared to rock. The theoretical performance potentials of commercial explosives are traditionally described by weight-strength (WS). Some explosive producers use the Langefors formula for weight-strength calculation, where also the gas volume is considered. Neither the theoretical nor the practical methods for determination of explosive performance such as the above, give good enough prediction of fragmentation and heave in rock blasting. For these reasons a project was started for the purpose of finding a small scale test method which predicts how different explosive parameters affect fragmentation and heave. Small scale tests were carried out by using 50 cm cubical blocks, mainly of concrete, but a few also of rock. All blocks were blasted in a steel cylinder with a diameter of 1.2 meters. The velocity of detonation was monitored, and the velocity of fragments was derived from their time of arrival at the steel wall, which was registered by accelerometers. After blasting all fragments were screened. The charge diameter was 22 mm and the amount of explosive 78 grammes, resulting in a specific charge of 0.62 kg/m3. Explosive properties such as energy, gas, VOD and density were systematically changed.The results presented in this paper show how fragmentation and heave systematically change when the explosive properties like energy, gas, VOD and density are systematically changed. The results from the small scale tests are compared to full scale tests carried out in tunnel blasting, and during summer 1990 also in bench blasting.

Jan 1, 1990

By A Batnasan

The main techniques for removing phosphorus as an impurity element in iron ores are physical separation (eg magnetic separation, flotation) and chemical separation (acid leaching, alkaline leaching and bioleaching) processes. When phosphorus occurs in iron ores as phosphate minerals (eg apatite), physical separation techniques are available to remove the gangue minerals from iron ores. On the other hand, chemical separation techniques are suitable for reducing phosphorus levels when phosphorus is associated with iron oxides and hydroxide minerals (eg hematite and goethite) in iron ores. In this study, the dephosphorisation behaviour of high‐phosphorus iron ore was investigated using a combined alkaline roasting and water washing route. Roasting experiments were conducted in a laboratory electric muffle furnace using various concentrations of sodium hydroxide (0–50 g/kg-ore) at different temperatures (100–800°C), varying prolonged roasting times (0–2 h), and different particle size fractions (-45 – -250 μm) of iron ore. Subsequent water washing experiments were employed at various pulp densities (30–1000 g/L) and different times (30 s – 60 min) at ambient temperature. Hematite and goethite were the main minerals in the high-phosphorus iron ore identified using X-ray diffraction (XRD) analysis. The iron ore assaying 63.40 mass % Fe (total), 0.240 mass % P, 4.53 mass % SiO2, and 4.87 mass % Al2O3 were determined using X-ray fluorescence (XRF) analysis. The results showed that the dephosphorisation rate of the iron ore was reached 40 per cent under roasting with 50 g/kg-ore NaOH at 300°C for 2 hours followed by water washing for 30 s with a pulp density of 300 g/L, reducing phosphorus level from 0.240 mass % to 0.142 mass %. At the same time, the iron grade was reduced from 63.4 per cent to 59.5 per cent, caused by an iron loss. Results suggest that the phase transformation of goethite induces the dephosphorisation of high-phosphorus iron ore under alkaline roasting conditions

Aug 22, 2022

A research and testing facility has been developed at Simtars to allow the spontaneous combustion testing of coal on a large scale. In this adiabatic reactor, the coal particle sizing is closer to that which could be found in the goaf and therefore the reaction is more likely to simulate the underground situation. This allows aspects of the spontaneous combustion process and methods for detecting a heating to be investigated which cannot be done in an operational mine. Up to 18 tonnes of run of mine or crushed coal is loaded into the reactor and allowed to self-heat while the temperature and gas profiles within the coal pile are monitored. The gases evolved from the coal are analysed using a fixed gas monitor system and gas chromatography to determine which of the common mine fire indicators and ratios can be used to predict the onset and progress of coal heatings. Benchmarking of the inherent propensity of a coal to spontaneously combust at both the large- and laboratory-scale is also possible by comparing the sample against previously tested samples. It has been found that the coal particle size affects both the time it takes for the coal to react and the location of the developing heating. The heatings generate approximately one metre from the inlet but in the crushed coal case do not progress to an advanced heating stage until the æhot spotÆ migrates towards the coal surface where there is an increased oxygen supply. Physical properties such as the compressive strength of the coal also affect the heating size and the time it takes for a heating to develop. Higher strength coals appear to oxidise more slowly and therefore take longer to self-heat as oxidation of the coal is the heat-generating step in the spontaneous combustion process. The best indicators of the state of the heating were found to be CO make and GrahamÆs ratio which are independent of air flow. Investigation of a potential spontaneous combustion should begin as soon as the CO make or GrahamÆs ratio exceed the background levels. The results from the tests undertaken on a number of Queensland and NSW coals are reported here together with the implications for the mining industry.

Jan 1, 2005

By J Rostami, T Afshar, M Sharifzadeh

Accurate estimation of rock load and selection of types of ground reinforcement to match post-failure behaviour of rock mass are important issues to assure tunnel stability. This paper focuses on the main factors affecting rock load to evaluate the capabilities of the existing models for prediction of rock load. Rock mass properties, project features and excavation method, as well as ground conditions and stress condition were found to be the key factors affecting the magnitude of rock load. Field data from tunnels of about 37.5?km driven in several geological structures in non-squeezing ground was used in this study. An in depth statistical analysis was performed on data to find correlations between the input parameters and measured/observed rock load. A weighting system was proposed to obtain the proper contribution of each factor for estimation of rock load in tunnelling through various ground conditions. The new classification system offered predicted values for rock load that were in a good agreement with the actual field measurements. The high disturbance factor (D>0.4) was the prime factor which controlled the rock load. Rock mass quality was the second key factor influencing rock load, which could be estimated by GSI, GSIresidual , especially when combined with tunnel depth.CITATION:Afshar, T, Sharifzadeh, M and Rostami, J, 2014. Estimation of rock load for tunnels with different disturbance factors, in Proceedings 15th Australasian Tunnelling Conference 2014 , pp 3–8 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 17, 2014

By Mercer J. K

Shaft guide reaetions in No. 2 Ore Shaft, Mount Isa Mines Limited, have been reduced to a low order of magnitude by the use of self aligning slippers and spring loaded guide rollers. Before commissioning the new 4500 h.p. winder at No. 2 Ore Shaft in January, 1958, it was realised that the increase in hoisting speed from 1800 to 2650 ft/min. would cause high shaft guide reactions which would be increased further when the elongated 12 ton skips replaced the existing 7 ton skips.The Cambridge accelerometer was used to measure the shaft guide reactions and to confirm that progressive skip modifications were resulting in reduced reactions. The skip modifications were completed in 4 weeks.INTRODUCTIONIn 1956 an expansion programme was initiated to increase production from 5000 to 11000 tons per mining day. At No. 2 Ore Shaft the 1500 h.p. winder operated at 1800 ft./min. and hoisted a 7 ton capacity skip. The new 4500 h.p. winder operates at 2650 ft./min. and hoists a 12 ton capacity skip. No. 2 Ore shaft is equipped with two 5 ft. by 5 ft. hoisting compartments (Fig. 1) and any increase in skip capacity meant an abnormally long skip. This feature, coupled with the increase in hoisting speed, could have caused very high shaft guide reactions. Before commissioning the 4500 h.p. winder, a Cambridge accelerometer was borrowed from the University of Adelaide and was used to measure the shaft guide reactions. Rather than carry out an extensive shaft timbering programme, the skips were modified by fitting cushioned guide shoes and spring loaded guide rollers (Fig. 1). Testing commenced in February, 1958, and was completed in March, 1958.

Jan 1, 1961

By G Harbort, K Quan

Traditional engineering practice for concentrator design uses a mixture of risk minimisation options and capital minimisation approaches. Often this is based on the preferences and experiences of the engineers involved.A geometallurgical approach involves identifying primary lithology, alteration, and mineral dispersion patterns within a deposit to allow representative samples for test programs. This provides a method to infer from laboratory test results meaningful production data for larger amounts of the ore body. Geometallurgical models can be integrated into the resource block model and subsequent mine plan and can assist in predicting when difficulties in the processing route might be expected, allow adjustment of the design to compensate for these or provide an early warning system so that production forecasts can be more adequately developed. Simulation modelling is the process of creating a computer model of a real system based on mathematical logic. In simulated time, the model explores the randomness and chronological sequencing of events observed on real systems. The objective is to mimic the real system with sufficient accuracy to make predictions of real behaviour under experimental conditions. Amec Foster Wheeler often use the ‘discrete-event’ simulation modelling technique. This method is well suited for inclusion of the statistical variation of effects in unit operation, mechanical availability, throughput and operating strategy (based on historical operating data for equipment downtime), operating schedules and equipment capacity. The net result of the simulation model is an analysis that can quantify risks in terms of system performance and capital and operating costs.CITATION:Harbort, G and Quan, K, 2017. Concentrator design – standard engineering practice versus geometallurgy and simulation, in Proceedings MetPlant 2017, pp 195–206 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 11, 2017

By Fawcett M

Pine Creek Goldfields Limited (PCG) is rehabilitating the Enterprise Pit to an aquatic habitat and potential water resource. This paper discusses the development and implementation of that plan. Pine Creek Goldfields Ltd's Enterprise Mine is located approximately 230 km south of Darwin. Open cut mining of the orebody commenced in 1985 and continued for eight years with 14 million m3 of ore and overburden extracted. Geochemical testing of exposed rock types in the pit has demonstrated a propensity to produce acid, and dissolved metals. A plan of management has been developed and implemented to minimise the risk of acid formation and improve the overall water quality in the final void. In developing the plan prime consideration was given to reducing the potential for pit water to impact upon the regional groundwater. An interception system was constructed on Pine Creek to divert approximately 9 km2 of upper catchment into the northern end of the Pit. Construction of the diversion system was completed in December 1993 and a comprehensive surface and groundwater monitoring program initiated. The monitoring program and associated studies are also discussed.

Jan 1, 1994

By G J JAMESON

The flotation cell has undergone a number of changes since the process was first introduced in Broken Hill in the early years of this century. The development of the mechanical flotation cell is reviewed, followed by a discussion of more recent developments such as the flotation column, the air-sparged hydrocyclone, the Wemco-Leeds cell, the Bahr and Deister columns and the Jameson cell.

Jan 1, 1992

By Faidel VV, Yushina TI, Yagodkina NG, Yatsimirsky AK, Chanturiya VA

Among the problems of the flotation of sulphides by the thiol collectors is the dependence and varieties of the flotation starting with the time and oxygen contents in the slurry. For sulphides, the flotation stage follows after the electrocatalytic oxidation of the collector with oxygen on the surface of the conductor minerals. The necessity of the oxygen for sulphide flotation (Plaksin and Bessonov, 1957) is in accordance with a corrosion model of a collector surface sorption (Salamy and Nixon, 1954). As Golikov and Nagimyak (1961) showed, the sulphides possess catalytic activity which provides the autoxidation of the xanthates in the sluny to the dixanthogen. This stage can be rather protracted, which limits the speed of the total process. Disulphides constitute important collectors for the sulphide flotation and the optimal degree of the xanthate oxidation is about 20 per cent for different ores (Abramov, 1978) or, following Chanturiya and Nazarova (1977) the surface compositions for various sulphide ores are different and an optimal oxidation degree is variable from five to 25 per cent according to the ore nature. The optimisation of the surface composition is possible by application of the dixanthogens, or by a partial chemical, electrochemical, or catalytic oxidation of xanthates (Desains, 1847; Schall, 1896; Vigdergauz, 1989). According to Rand (1977) and Vigdergauz (1991) sulphides possess different catalytic activities for oxygen electroreduction. In neutral solutions, for example, the following decrease in the order: FeS2 (Fe,Ni)9S8 CuFeS2 FeS; Cu5FeS4 Cu2S; PbS (Vigdergauz, 1991). For the cases of low catalytic activities of the sulphides, the preflotation oxidation of xanthate could be useful to optimise the process kinetics.

Jan 1, 1993

By K McSweeney, A Marton

This paper discusses the development of a strategy for ensuring long-term geotechnical stability of final pit walls at AngloGold AustraliaÆs Sunrise Dam Gold Mine. The strategy makes use of improved blasting techniques, changes to the mining sequence and other geotechnical techniques. Only the blasting techniques will be discussed in this paper. These were developed from numerous trials over a period of several months from May 2003 through to November 2003. The focus was to reduce crest loss damage and thereby improve final wall integrity, maintain designed safe berm widths, and ensure long-term stability. Crest loss mechanisms are now understood and are explained in this paper. The preferred solution is outlined; which is a combination of improved blasting and changes in the mining sequence to cater for pre-existing ground conditions. These ground conditions include dynamic hyper saline water with sometimes intersecting steeply dipping discrete structures. Project requirements were defined as mitigating or eliminating crest loss using 127 mm diameter pre-splitting and targeting minimal production blasting vibration levels against the walls. Blast results were measured by post blast visual inspections and video. Blast design input utilised the predictive algorithms built into simulation software when attempting to minimise vibration levels against the walls. Protection of the already fired pre-split line is now achieved using small charges against the wall and timing sequences which form a mask for the later firing main charges. The pre-split spacing is now modified to suit the geology. There eight blast domains identified around the perimeter of the pit, each having varying joint orientation. In addition, blasting the bench adjacent to the berm crest has particular considerations not required by the middle and lower benches. The preferred solution was to purposefully translate this damaged top section of the pre-split bench into a planned subsequent dig zone. The design of this top æweightÆ is a direct response to the nature of the crest loss mechanism.

Jan 1, 2004

Tomago sandbeds is an unconfined coastal aquifer located 20km. north of Newcastle in New South Wales, Australia. These sandbeds extend over an area of about 130km. The average aquifer depth is about 20m. with a mean transmissivity of about 400sq.m/day and a specific yield of about 0.2. The city of Newcastle and surrounding areas derive water from Tomago for potable purposes. The average extraction being about 50 ML/day. The iron value of the ground water varies between 5 and 15mg/1. This iron is removed by a conventional water treatment plant using Alum. The aquifer whilst principally quartzose also contains deposits of heavy minerals like rutile, zircon, ilmenite, etc., The mineral sand mining has been in progress at Tomago since 1972. Mining is carried out in a con- structed pond with a water depth of about 5m. on which is floated a suction cutter dredge that moves across the front of the pond using a water jet to entrain the soil from the ground surface to the bottom of the mineralised zone. Effect of mineral sand mining on ground water quality, particularly iron, was the subject of intensive investigations undertaken at Tomago since 1970. The effect on water quality was studied by collecting and analysing water samples from network of observation wells established during the period of mining. Water samples were collected about 4-5 years prior to mining and analysed for iron, sulphate, chloride, pH, etc., After mining the observation wells were re-established at the same locations and water samples were collected and analysed for at least 5 years after mining. Water samples collected from majority of the observation wells indicate an increase in iron levels. The level of increase is very site specific. The increase in iron is also

Jan 1, 1987

By N McIntyre, J Smart, J Ossa-Moreno, D Rivera

Hydro-economics is a field of research that analyses water resources from a multi-disciplinary approach (Harou et al., 2009). These models generally include one component that describes the water resources of the catchment being analysed (e.g. the physical behaviour of water), an economic component that describes users’ water demand function (changes of willingness to pay relative to availability) and a coupling methodology. These models go beyond calculating cash flows for water investments and analyse the economic behaviour of users when facing water shortages or surplus. In other words, water demand curves are defined for each water user (e.g. urban, agriculture, mining, environmental and hydropower), and they are used to retrofit the water resources component. As a result, the model approximates the catchment scale response to droughts and wet periods, including the effects of water markets, restrictions, recycling, water sharing and any cooperative efforts by users to adjust to changing climate conditions.These models are usually focused on agriculture and urban users, the main water consumers in most catchments, but they have sometimes included environmental, industrial and hydropower users (Bekchanov, Sood, & Jeuland, 2015). Nevertheless, although a relevant user of water in several catchments, the mining industry has been rarely included. This paper is part of an ongoing project that addresses this gap by developing a hydro-economic model in the upper section of the Aconcagua River in Chile. The aim of the paper is to analyse the sensitivity of the outputs of the model to changes in input data, both hydrological and economic, and to develop insight on usefulness of such models for reducing water conflicts in mining regions.CITATION:Ossa-Moreno, J, McIntyre, N, Smart, J and Rivera, D, 2018. Sensitivity of a hydro-economic model in a mining region to changes in hydrological and economic inputs, in Proceedings Life-of-Mine 2018, pp 52–55 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 25, 2018

By M Binks

The Fosterville Gold mine is located approximately 20km to the East of Bendigo in central Victoria. The gold at Fosterville occurs as solid solution within disseminated arsenopyrite and pyrite. The Fosterville ore bodies contain various amounts of native carbon in the form of bituminous coal. This carbon (Non-Carbonate Carbon or NCC) has been the predominant mechanism for gold loss from the processing facility through”preg-robbing”.Processing of the Fosterville ore is achieved initially through crushing, grinding and flotation to extract the sulphides. The sulphide concentrate is oxidized using Bacterial Oxidation, before being leached in a conventional CIL circuit. A high portion of the native carbon (NCC) in the mine ore is naturally hydrophobic in nature, and subsequently reports to the flotation concentrate stream, and ultimately onto the CIL circuit. This NCC has a notable preg-robbing ability. Treatment of Black Shale ores, which have elevated NCC levels, has historically resulted in CIL recoveries as low as 35%, with around 60-80% of the gold loss from the leach circuit attributed to preg-robbing.Standard technologies for dealing with a carbonaceous leach feed have been trialed with limited success. The discovery of the significance of heat on leach recoveries, triggered extensive testwork on a range of leach feed and tails samples. Pilot plant testing demonstrated that “Heated Leaching” of the CIL tails was the best process with an average recovery increase of 7.5% being achieved. Following a successful pilot study, an engineering feasibility study was completed by Minerva Engineering, and the project economics determined. With strong recovery gains evident and a predicted project payback of 1-1.5 years, approval for the installation of the full scale plant was granted. Detailed design commenced in September 2008 with installation commencing in January 2009. The circuit was successfully commissioned in April 2009.Following commissioning, the Heated Leach circuit has achieved recovery gains of 4-14%, and has proven itself to be a significant contributor to the overall plant performance.

Aug 8, 2011

By D Pienaar, B McFadzean, C T. O’Connor

Collector and frother molecules in sulfide mineral flotation are thought to have separate roles in the flotation process, where the collector selectively hydrophobizes the mineral surface and the frother adsorbs onto the bubble surface, inhibiting bubble coalescence. Molecular interactions have been observed between these molecules which allow them to co-adsorb at each of the respective interfaces, improving pulp phase recovery. This study investigated the molecular interactions between sodium diethyl dithiophosphate (SDEDTP) and hexanol, and the mechanism by which these interactions improve the recovery of galena. In this study each of the respective interfaces were investigated for the adsorption of SDEDTP. The adsorption of SDEDTP on galena was measured by UV-Vis spectrophotometry and isothermal titration calorimetry. At the air-water interface foam stability and pulp bubble coalescence was used to quantitatively discern if there are interactions between SDEDTP and hexanol. Microflotation and induction time experiments were performed to determine the effect of SDEDTP, hexanol and their mixtures on mineral recovery and bubble-particle attachment. The results of all these tests were collated to determine if SDEDTP-hexanol mixtures synergistically improve sulfide mineral recovery. It was found that SDEDTP reduced bubble coalescence in the pulp demonstrating activity at the airwater interface. Together SDEDTP and hexanol synergistically enhanced foam stability. This indicated that there are attractive molecular interactions between these reagents. Even though DTP did not adsorb onto galena, DTP-frother mixtures synergistically improved the recovery and attachment probability of this mineral. Adsorption studies with aeration showed that DTP in the presence of a frother transfers from the air-water interface to the galena mineral-water interface during bubble-particle collision. It was proposed that the transfer of hexanol through the thin film enhanced the kinetics of film thinning between the bubble and the particle by destabilising the thin film thus improving pulp phase mineral recovery.

Aug 24, 2022

By M Szwec, P Corbett, P Sheffield

Centennial Coal operates the Springvale and Angus Place longwall mining operations that extract coal beneath the Newnes Plateau in the Western Coalfields of New South Wales. Springvale and Angus Place mines have a history of difficult geotechnical conditions, where mining operations are conducted with a thick coal roof interbedded with numerous claystone units and relatively high levels of vertical and horizontal stress. The combination of a relatively high stress regime and weak laminated roof has led to numerous roof failures throughout the mine’s life.Following underground roadway development, the mine roadways can experience high levels of roof movement that can lead to roof falls. Two-point wire extensometers (known as ‘telltales’) are extensively used for roadway condition monitoring. Along with extensive other geological and geotechnical data, monitoring data is used to prepare geotechnical hazard plans that are then used in all areas of mine planning to allow for safe and efficient operation of the mine.This paper presents a case study of the use of a new software tool that allows time-scaled visualisation of monitoring data in the context of longwall and development face positions and installed strata support. It can be used in the identification of geotechnical risk factors and their interaction with the mining process to allow a more robust geotechnical hazard process. It interfaces directly with the widely used ExtoChart database software and can be presented with scale drawings of geological/geotechnical data to allow correlation of monitoring data with risk factors.CITATION:Corbett, P, Sheffield, P and Szwec, M, 2014. A new tool for extensometer data analysis and improved understanding of geotechnical risk factors, in Proceedings AusRock 2014: Third Australasian Ground Control in Mining Conference , pp 217–232 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 5, 2014

By P Hagan, S Wang, D Li, B Loh

A study was undertaken to quantify the differences in performance between a full-length encapsulated cable bolt and a partially encapsulated cable bolt. Much research has been reported on the differences in cable bolt design and operational parameters such as borehole diameter and grout type with much of this based on the Laboratory Short Encapsulation Pull Test (LSEPT). However, there is a gap in knowledge surrounding the effect of anchorage length on load transfer and cable bolt performance particularly with respect to the role of the bulb in modified cable bolt designs. The objective of this study was to expand the existing knowledge base of cable bolt performance with the intent of developing a revised design criterion. Alongside a comprehensive review of current cable support theory, design and practice, a series of pull tests was undertaken to compare each anchorage method using a high capacity modified cable bolt. The results serve to confirm earlier work on fully encapsulated anchorage. Both the peak load capacity and stiffness of the fully encapsulated cable bolt were nearly double that of the partially anchored cable bolt. When the tests were repeated in an oversized borehole, while there tended to be more variation in performance, there was little significant difference in the mean results. In addition to anchorage length, there are other factors that contribute to cable bolt performance including in order of sensitivity: embedment length, water-to-cement ratio of the grout, type of grout and borehole diameter. CITATION:Loh, B, Hagan, P, Li, D and Wang, S, 2018. A comparison of the performance of fully versus partially encapsulated cable bolts based on laboratory-scale testing, in Proceedings The Fourth Australasian Ground Control in Mining Conference (AusRock), pp 291–302 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 30, 2018

By I Jabouri

There are many physical factors that can affect the self-heating rate of coal. The presence of seam gas has often been referred to as inhibiting coal self-heating due to the limited access of oxidation sites created by the presence of the gas adsorbed on the coal pores. Similarly, the presence of bed moisture in the coal also acts as an inhibitor of oxidation by blocking access of air into the pores. Gas drainage of a coal seam prior to mining removes both gas and moisture from the seam. Bulk coal self-heating tests in a two-metre column on both gassy, as-mined and gas-drained, dried high volatile bituminous coal show that removal of gas and moisture from the coal accelerates the rate of self-heating to thermal runaway from 8.5 days to 4.25 days, from a start temperature of 30¦C, with an airflow of 0.25 L/min. The corresponding gas evolution pattern for each of these situations is different. Therefore, it is necessary to take this change in coal condition into consideration when developing a spontaneous combustion management plan.

Jan 1, 2005

The problems of the as-built water supply system are discussed. Recent modifications to the collection dams and the water cleaning station using wedge wire screens to resolve these problems are outlined.

Jan 1, 1978