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By A G McLean

This paper will first discuss the fundamental requirements for reliable discharge of backfill by gravity alone at high per cent solids. This discussion will include brief mention of the underlying principles of saturated solids flow contrasted to that of stirred tank or dilute discharge systems. Here the need to possess full knowledge of the flow properties of the stored material will be highlighted. In regard to attainment of reliable discharge it will he revealed flow from a backfill silo will only occur provided: ò the flowing material in the converging channel is yielding throughout; ò slip is occurring along the walls; ò an arched stress field is present in the vicinity of the outlet; ò the flowing backfill must be dilating to flow in response to the dilating stress field; ò sufficient dilation water is injected, in the hopper, to eliminate the formation of adverse interstitial pressure gradients; ò sufficient dilution water is injected, downstream from the hopper outlet, to generate a pumpable slurry of maximum per cent solids; ò the silo level must be sufficient to prevent the pump suction forming a rathole; ò the system must be configured and operated to generate stable discharge density; and ò the system must be configured and operated to minimise damage to the pump liners. In addition proper account must be made of the silo overflow weir and settling capacity to match plant throughput. Here opportunity will be taken to highlight the complexity of this multi-phase flow situation as well as the general yet paramount requirements for reliable solids flow. It will be shown that by proper attention to the multi-phase considerations gravity discharge of slurries as a consolidated bed with minimal water content is possible. Such high solids discharge results in greatly reduced demand for process water. In addition pumping costs are reduced as is the delay time associated with the mine drainage operation. Confirmation of successful applications will be highlighted by reference to several recent case studies. In the concluding phase mention will be made of the key yet very simple and effective operation techniques which enhance reliable Discharge.

Jan 1, 1998

By D R. Andrew, K T. Morrison, J F. Tiedgen

Waste blasting in iron ore operations can pose a number of technical challenges, including blasting through variable geology and blasting soft ground in wet conditions. Blast patterns are designed to be loaded with a specific range of explosives with unique characteristics and properties. Each explosive product will produce an optimal blast outcome for a given set of blast parameters and ground conditions. Sites characterised by softer rock typically require low-density, low-energy explosives. Historically, these products have contained a high ratio of porous ammonium nitrate prill and are only suitable for dry blastholes. When blastholes within a pattern are found to be unsuitable for such explosives due to wet, damp or muddy blasthole conditions, a new product is selected. While it is easy to switch products to match blasthole conditions, the different properties of the new non-ideal product can impact blast performance and increase loading times and costs.Flexigel™ is a solids-sensitised emulsion blend suitable for loading at variable densities and energies in dry, damp, muddy or wet blastholes. It is a universal product that can be loaded to match ground and blasthole conditions or complement other explosive products.At Fortescue Metals Group’s Cloudbreak site, Flexigel™ was trialled through a series of matched pair blasts in order to optimise blast outcomes in conditions that were non-ideal for existing explosive products and to reduce the overall cost of drill and blast operations. Flexigel™ yielded improved instantaneous dig rates and improved loading efficiency when compared to a traditional emulsion blend. The results also indicated that variable-density and variable-energy products afford greater flexibility for optimising blast outcomes across all ground and blasthole conditions trialled.Trials of Flexigel™ at Cloudbreak have shown that a variable-density solution is available for conditions where traditional explosive products are non-ideal. As such, Flexigel™ is able to fill a major void at any operation where the desired low-density, low-energy explosive product must be substituted for a more expensive traditional emulsion blend or heavy ammonium nitrate/fuel oil.CITATION:Andrew, D R, Morrison, K T and Tiedgen, J F, 2015. Flexigel™ – a variable-density solution, in Proceedings Iron Ore 2015, pp 413–418 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 13, 2015

By D Piason

Newstan Mine commenced mining operations in 1887 and has since undertaken extensive mining within the Young Wallsend, Great Northern, Fassifern, Borehole and West Borehole coal seams using a range of mining methods including board and pillar mining and longwall mining.Newstan Mine is planning to extend the existing mine workings using longwall and miniwall mining methods to produce up to 4.5 Mt/a of run-of-mine (ROM) coal for at least 25 years. The planned extension in the Main East and South West sections will involve developing new longwall panels and related mains headings. The new area contains higher specific gas emissions (SGE) than previously experienced at Newstan Mine. To achieve the optimum ventilation requirements for the longwall operation in this new area, a new ventilation shaft was designed to be constructed using the blind boring method in the South West Headings section.This paper presents the shaft design parameters and selection criteria for the construction of the new ventilation shaft, including the construction phase and the holing process to connect the shaft to the underground workings. The paper also discusses future ventilation requirements for Newstan Mine.CITATION:Piason, D, 2015. South-west ventilation shaft project at Newstan Mine, in Proceedings The Australian Mine Ventilation Conference, pp 229–238 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 31, 2015

MIM Holdings Limited (MIM) committed to a long-term future in copper in the Carpentaria Mineral Province in 1996 with the development of the Enterprise and Ernest Henry mines, and an upgrade of MIMsÆ copper smelter in Mount Isa, and copper refinery in Townsville. The upgrade of the copper smelter will enable the smelter to produce 250 000 tpa of copper with a single primary smelting unit, the copper Isasmelt furnace installed in 1992. The copper Isasmelt furnace currently smelts concentrate containing 145 000 tpa copper. To reduce the risk associated with the copper smelter upgrade, and shorten the ramp up time of the upgraded smelter, two high-rate trials were undertaken in 1997 to demonstrate that the copper Isasmelt furnace could, with an additional 525 tpd of tonnage oxygen, treat concentrates containing 250 000 tpa copper. During the trials the concentrate feed rate to the Isasmelt furnace was increased from its average rate of 100 tph of concentrate to 150 tph of concentrate. The increase in the capacity of the furnace was achieved by: supplementing the installed tonnage oxygen capacity of 595 tpd with an additional 300 tpd supply; reducing variability in the furnace feed rate from ¦25 per cent to less than ¦5 per cent; and reducing tramp air. The Isasmelt furnace smelted concentrate at a rate of 140 to 150 tph for four hours 25 minutes during Trial 1, and at a rate of 150 tph of concentrate for five hours and 30 minutes during Trial 2. The oxygen utilisation was 100 per cent during the trials with the process remaining under control. As a result of the development work for the trials the average treatment rate of the furnace under normal conditions increased by ten per cent.

Jan 1, 1998

By J Player, R de Vries, J F. Carswell, R Hassell

"The feasibility of Dugald River mine was highly sensitive to mining and geotechnical design assumptions, particularly stope dilution and strike length. Following the start of initial mine development in 2012 the project was suspended in 2013 due to concerns over the complex ground conditions and resulting dilution, stope sizes and general mineability. A technical risk mitigation program was undertaken with a key aim to increase the level of confidence in predicting stope dilution.A stope sizing and dilution model was developed that provided a range of stope sizes based on geotechnical conditions and acceptable dilution. The original feasibility used a standard stope size of 25 m height by 25 m span. The new model was used as the foundation for the subsequent updated development plan for Dugald River that reduced much of the previous technical risk through greater confidence in the geotechnical and mining parameters.Several methods to predict stope hanging wall dilution and appropriate stope size have been used at Dugald River that relied on the modified stability graph method. Improvements to the method were identified as more rock mass and mining data became available through the completion of a trial stoping program, increased underground development and a large geotechnical core logging campaign. These improvements not only focused on the technical aspects of the model, but also on creating an easy to use and sustainable process that could be used at multiple stages of mine planning from life-of-mine (LOM) through to final stope design.Several of the main improvements are: understanding stope failure mechanisms and collecting only geological information that is most relevant detailed modelling restricted to the rock mass that could be expected to influence stope behaviour accounting for different stope failure mechanisms with the stability chart method assigning a geotechnical confidence factor to the data making it useful as a mine planning tool with integration into the mine planning process using performance of completed stopes to update and calibrate the model.The model provides recommended hydraulic radius and associated dilution estimates across the ore body. Providing the hydraulic radius allows planning engineers the option to consider single, double and triple lift stoping options suitable to mine scheduling. While the model has been calibrated for conditions at Dugald River the process used to create the model is not site specific and is applicable to other steeply dipping tabular orebodies.CITATION:Hassell, R, de Vries, R, Player, J and Carswell, J F, 2017. Improving the prediction of stope hanging wall dilution – a critical factor to the success of Dugald River, in Proceedings 13th AusIMM Underground Operators’ Conference 2017, pp 71–78 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Oct 16, 2017

The implementation, during recent years, of the concept of distributing chilled water to improve cooling at the working faces represents a very important innovation for the mining world. In a logical extension to that concept it has been recognized that distribution of the cooling water at high pressure would provide a potential source of energy for driving water powered hydraulic machinery such as rockdrills, water jets for moving rock, and support props, in the workings. With this in mind, the Research Organization of the Chamber of Mines of South Africa is pursuing the development of such a dual system that will provide cooling in a deep mine and, at the same time, drive hydraulic machinery. A comprehensive design of such a system has been completed and has shown the concept to be entirely feasible technically using technologies currently available. A pilot system is working in a mine shaft 2000 m deep giving a pressure of 20 MPa at the bottom of the shaft and distributing the cold water into the workings at a pressure of 17 MPa. The system shows considerable potential for improving existing mining methods and will open the way for the development of new mining methods for deep mining operations.

Jan 1, 1986

WaihiÆs Cornish Pumphouse is located on the edge of town in a dynamic and changeable area where history affects the future. As in any successful project, ingenuity and team work have been significant factors in the planning and implementation of the buildingÆs relocation.   Work around the Pumphouse site commenced circa 1900 with preparation to sink the No 5 shaft. Construction of the pumphouse and the steam driven C pump was completed by 1904 with hints of what was to come evident in 1905 as the No5 Shaft collar had sunk 350mm in four years. By 1907 surface cracking developed over the Milking Cow cave area and in 1913 200 mm was cut off the main pumping rod as the ground was sinking at a rate of 25 mm per year.   In 1913 ground around No 6 shaft had subsided 600 mm.It was also the end of an era as steam gave way to hydroelectricity and electric pumps replaced the C Pump. C Pump was maintained as a backup from 1913 to 1929 and then finally scrapped. 25 years of underground caving operations removed approximately 400 000 cubic metres of ore leaving a 7 ha collapse crater that eventually became known as the mine lake.   On the edge of town a tension zone over the Royal Lode began to develop 50 m diameter chimney collapses as large scale blocks relaxed into unfilled stopes in the old Martha underground mine. After the 1999 collapse part of the main street was closed and after the 2001 collapse Hazard Zones were declared, more streets closed and areas evacuated and retired from residential use.   By 1984 the Pumphouse has approximately one degree of tilt and in 2002 tilt on the Pumphouse accelerated briefly and then settled down. Tilt accelerated again late in 2004 and a scoping study was initiated to define protection options for the pumphouse. The Martha Mining Licence requires Newmont to 'Take whatever precautions are necessary to protect the old Cornish Pumphouse adjacent to No5 shaft' and viable solutions needed to found.   Any work on the Pumphouse comes standard with a wide range of stakeholders, governing agencies and legal framework to consider. Land Information New Zealand (LINZ) is owner and administrator of the pumphouse.

Jan 1, 2006

By Hein K. A A

The Mt Todd Gold Project is located 42 km northwest of the township of Katherine within the Northern Territory of Australia. It is situated in the southeastern portion of the Pine Creek Geosyncline. The Batman gold deposit, the largest of several deposits at Mt Todd, hosts an identified mineral resource of 92.9 million tonnes (Mt) grading 1.4 grams/tonne (g/t) as at 31 December 1993. Host rocks at Batman are assigned to the Early Proterozoic Burrell Creek Formation, comprising greywacke and siltstone interbedded with shale. Gold mineralisation is associated with a very broad zone of sheeted and stockwork quartz-sulphide veins that occupy a corridor of intense brittle fracture and faulting. A conceptual model of formation involves the dilation of joints and micro-faults within overlying sediments during sinistral movement along a basement shear zone. Synchronous granite intrusion generated convective circulation of crustal fluids through the sedimentary pile resulting in the formation of auriferous quartz veins within favourable structures. The Batman zone of mineralisation is not directly related to fold structures which typify several other major gold deposits within the Pine Creek Geosyncline.

Jan 1, 1994

By D Adhikary, M Khanal, B Poulsen, H Guo

During longwall mining, the ability to reliably estimate the height of longwall mining-induced fractures is critically important in order to effectively manage mining-induced water inflows and aquifer interference. The extent of mining-induced subsurface deformation, fracture, surface subsidence and aquifer interference is largely controlled by local lithological conditions, mining methods and mine layouts. Therefore for the purpose of quantifying mining impacts it is imperative to incorporate the local geology and mine plan parameters in a mine scale model. Two different numerical models based on 3D continuum (Finite Element Method) and 2D discrete (Discrete element method) domains, respectively, are developed to incorporate the mine site parameters of a study site in Australia. Both the models are calibrated and validated with measured mine data. The continuum numerical modelling method is used to investigate the change in permeability of the overburden strata due to underground extraction considering the full 3D nature of geology and mine layout. The discrete element modelling method is used to simulate the initiation and propagation of fractures, allowing the caving of overburden strata to be represented in a realistic manner, and the permeability enhancement of the strata overlying the goaf to be calculated from first principles. Permeability enhancement in the discrete element model is a direct function of the fractures induced in the deformation/caving process; fracture location, extent, aperture and connectedness. From the numerical modelling under the simulated condition for the mine studied here, it is predicted that the permeability of the caving zone from the goaf to five times the mining height increases by almost six orders of magnitude. In the fracture zone to approximately ten times the mining height, the change in permeability is predicted to be around two to three orders of magnitude. At the height of more than ten times the mining height the magnitude of permeability increased is predicted to be one to two orders of magnitude. CITATION:Khanal, M, Poulsen, B, Adhikary, D and Guo, H, 2018. Modelling of height of connective fractures during longwall mining, in Proceedings The Fourth Australasian Ground Control in Mining Conference (AusRock), pp 254–265 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 30, 2018

In this year marking the centennial of the Geological survey of Western Australia, it is appropriate that we should review the value of the contribution of the Survey, and Australian Geological Surveys in general, to the exploration industry. Most geological surveys in Europe and North America had their origins in the first half of the 19th Century when the burgeoning Industrial Revolution was demanding the identification of new sources of coal and iron ore. Australia, as a young nation, essentially dependent on an agricultural and pastoral industry, and with already discovered vast coal resources in Eastern Australia, had, no great need to explore for new mineral resources. However, with the discovery of gold in 1851, the mineral potential of the country became apparent to the whole world and the first Geological Survey in Australia was set up in 1852 in Victoria under A.R.C. Selwyn. After a number of false starts all Australian States had an established Geological Survey by 1900. It is hard to assess the contemporary value of the early work of the Geological Surveys with respect to exploration, almost all major mineral discoveries were made by prospectors and most mining companies had little time for geologists - although the Chamber of Mines in Western Australia in 1910, did recommend that the Geological survey be granted an additional ten thousand pounds over two years to map all existing goldfields and the country between them, and produce a reconnaissance map of the area east of a line between Wiluna and Edjudina as far as the South Australian border!

Jan 1, 1988

By A J. Ryan

The Ernest Henry copper-gold Mineral Resource is the major asset heldby Ernest Henry Mining Pty Ltd, comprising A$6.4 billion of containedmetal. The case history presented below describing the evolution andevaluation of this resource generally observes best industry practice and is a good example of effective due diligence in the resource delineation and estimation process. Early reconciliation data from mine production confirms the accuracy and validity of the resource estimate, providing valuable assurance to the mine owners.

Jan 1, 1998

By McMahon DW

The major components of a mineral reserve estimate for project development purposes are described with respect to sampling, geology, estimation methods, mining factors, metallurgical factors, environmental factors, financial and other factors. Each of these components impact to varying degrees on the recoverable metal or product of a new mining project. The evaluation of a project is a complex process involving the best the profession can provide in terms of geological understanding, engineering ingenuity, metallurgical skills and financial evaluation.

Jan 1, 1983

The major coalfields of Australia are almost unique in the occurrence in some areas of CO partly or almost completely replacing CH4,t1e normal seam gas from metamorphism of coal due to time and depth of burial. This CO2 situation has some benefits but mostly it introduces particular safety problems result- ing from the high density of CO2 relative to air, magnifying problems of gas dilution in ventilation, and the greater affinity which coal has for CO2 than for CH4 increasing proneness to gas-dynamic phenomena. The occurrence of most CO2 seam gas can be directly related to igneous intrusion and to its associated pneumatolytic injections. Whilst intrusions locally increase the rank of coal by their contact metamorphic effect, normal coal metamorphism due to time and depth of burial continues after intrusion and injection together with the ongoing metamorph- ic production of CH . This new CH slowly displaces any CO to leave some present-day intruded coals with seam gas with reduced CO2 or even without any significant CO2 associat- ions. The seam gas maps of the Bowen and Sydney Basins and more detailed maps of specific areas exemplify these situations. Mining progress, plus more searching studies of igneous-coal relationships provide improved understanding of the factors involved.

Jan 1, 1986

The Australian Mineral Industry's growth in the last two decades is directly related to the skill and experience of its consultants in mining, metallurgy, geology, finance, management, marketing, and engineering. The role of mining consultants needs to be re-examined, so that the quality of the service they deliver will continue to be adequate for the 1990's. One means of this evaluation is by examining Case Studies in the strategies and difficulties of developing contract and entrepreneurial research.

Jan 1, 1986

Olympic Dam Operation is located in the Stuart Shelf Geological Province approximately 520 km north-north-west of Adelaide in South Australia. It is estimated to be one of the worldÆs top five copper orebodies by size and the biggest single occurrence of uranium. The deposit was discovered by Western Mining Corporation in 1975. Exploration was based on a conceptual model for the formation of sediment hosted copper deposits together with gravity and magnetic anomalies.The capital cost of the current expansion is estimated to be A$ 1.64 billion. The underground expansion includes a total of 36 km of lateral development, a railway haulage system delivering ore from a series of ore passes to a new crusher station and a significant increase in ventilation capacity and mine services.In the current underground expansion the mining method will essentially remain unchanged. The ore handling system will incorporate some recent advances in technology, improving efficiency and safety standards.Major infrastructure excavations are designed by a team of engineers attached to the Olympic Dam Expansion Project and a number of consulting groups. Once the expansion is fully completed it is expected to increase the current annual underground production level from 3.4 Mt to 9.0 Mt of ore by the year 2000. Mine design engineers working for the Olympic Dam Operation have recently completed a three year final design plan to achieve this target.Access to the existing mine is through the decline and Whenan shaft. The Whenan shaft was originally sunk as an exploration access and later upgraded for hoisting. The 4 km long service decline connecting to the Whenan shaft was later constructed to accommodate the increase in service demand. The Robinson and Sir Lindsay Clark shafts were developed at later stages to cater for the hoisting during and following the expansion lead up.

Jan 1, 1999

By H Kusumi, A Nishio, K D. Halim, Y Yun

In Japan, evaluation of rock mass is crucial in mountain tunnels that are constructed using the NATM (New Austrian Tunneling Method) to determine the support pattern that in turns determine the final support structure of the tunnel. The current evaluation method is highly dependent on the observations of on-site expert engineers that grades the rock mass in four levels across nine criteria: A. Condition of tunnel face. B. Condition of excavated surface. C. Compressive rock strength. D. Weathering. E. Spacing of discontinuity. F. Condition of discontinuity. G. Orientation of discontinuity. H. Spring water. I. Water degradation. However, in recent years, the number of said experts are decreasing rapidly due to labour shortage caused mainly by an aging population and declining birth rates. This is one of the biggest challenges that the construction sector currently faces, and this may affect the efficiency of tunnel construction projects in the future. Thus, this study proposes the implementation of a deep learning tool called the convolutional neural network (CNN) to quantitatively evaluate rock fractures on tunnel face. In addition, Gradient-weighted Class Activation Mapping (Grad-CAM) is applied to visualise CNN, as well as verify the applicability of CNN to the evaluation of rock fractures. In other words, by creating heat maps of tunnel face images, CNN is utilised to evaluate rock mass in categories that are associated with discontinuities. The CNN model is most accurate when six convolution layers, four pooling layers, and three dense layers are used. This yields an accuracy of 75.0~88.0 per cent. Thus, the CNN model is verified to be feasible method to evaluate rock fractures.

Nov 29, 2022

By B Andrew, L Krois

Ground control systems, including ground control management, play an integral role in the effective and efficient extraction of copper ore at the Xstrata Copper û Mount Isa Copper Operations (MICO). Rock mechanics engineers and support technicians at Xstrata Copper have undertaken a wide range of investigations and actions over the past few years. Outcomes from this work have resulted in the evolution of ground control management practices on the site. Specific areas of improvement include the installation of efficient one pass ground control systems, greater application of sprayed concrete for areas requiring rehabilitation and the addition of new improved products, the removal of obsolete and less efficient products used in standard ground control practice. In some areas underground, the most effective methods to manage ground control hazards continue to be a variation on proven methods that have been in use for many years in the MICO. Ground control systems for developing through the basement contact zone are very similar to that developed in the late 1980s. Similarly, ground control systems have gone through repetitive evolution, whereby some systems fall in and out of favour subject to perceived costs. For example the pouring of concrete shuttering around drawpoints in poor ground conditions has recently been re-accepted as an effective means to manage this poor environment û despite having been previously cast aside as too expensive. Ground control management systems continue to evolve towards the goal of creating a workplace that is free of rockfall-related injuries and provide for higher levels of efficient and effective techniques by which ore is extracted from the ground.

Jan 1, 2005

By M Gwynne

McArthur River Mine is located approximately 50 km south-east of the NT township of Borroloola and approximately 120 km from the Gulf of Carpentaria. The mine hosts a series of layered shallow dipping tabular orebodies, with current reserves of approximately 22 Mt, grading 13.6 per cent zinc, 6.3 per cent lead and 65 g/t silver. Owing to economic considerations, the lower most orebody (No 2 orebody) has been selected for initial mine production. Following the completion of the No 2 orebody, mining will transfer to the lower grade and metallurgically more complex No 4 orebody. The mine employs a total of 235 people throughout the underground and surface areas. The majority of the workforce work a fly-in/fly-out, seven-day- on/seven- day- off roster with 12-hour shifts. Presently, a further 18 contractors are involved in trial mining areas. An apparent dip room and pillar technique has been selected as the primary mining method. Officially opened in September 1995, the current production rate of 1.2 Mt per annum will eventually escalate to 1.5 Mt per annum. Though most production is achieved through the use of conventional loaders, trucks and jumbos, the recent development of methods utilising scrapers, hand held machines and modified jumbos has produced encouraging results in steeper dipping areas of the mine. Initial production has been hindered by several factors including extensive and largely unpredictable faulting, worse than expected ground conditions and difficulties with some mining equipment. Compounded by the relatively thin 3 m ore seam, initial mining resulted in high dilution and poor production. This paper describes the development of unique planning and production systems that were evolved in order to overcome these problems and achieve the targeted production rates and head grade.

Jan 1, 1998

The purpose of this paper is to look at the costs of tailings disposal, and in particular the comparative costs of different disposal methods in different situations. Costs are often the hidden design parameter. We know that they are important but sometimes feel that it is not very respectable to talk about them in public. It is better from a public relations point of view for it to appear that the decision-making process is driven by environmental factors or best practice rather than looking for the cheapest option. But this presumes that the two are mutually exclusive and this need not be so. These issues are discussed later in the paper. It is also commonplace for the scale of the mining operation to change with time, typically involving increased throughput rates and total tonnage. New ideas in tailings disposal make their appearance. The regulatory attitude can change. Having made a decision about the disposal method at the time of original feasibility study and design it is easy to suppose that the mine is æstuck with itÆ. But that also need not necessarily be so. Careful costing of options may reveal that a switch in method could be cheaper in the long run. The designer will often tend to think in terms of initial earthworks and site development costs. Aspects of power consumption, labour, and periodic campaigns to move tailings pipelines and carry out embankment raises are either overlooked or not given the same emphasis. This can distort the evaluation process. It is productive to think in terms of dollars or cents per tonne as a basis of comparison, not just the total sums. Tailings disposal can cost from as little as 10 - 20 ó/t to over $1.00/t which is a big range. There may be good and unavoidable reasons for the high figure but there would need to be to justify being so far over the industry average. At the same time, mining and processing costs, at least for gold, are around $15 - 20/t. Viewed from this perspective tailings disposal will always be a relatively small component of the annual budget.

Jan 1, 2000

By R Castedo, P Segarra, L M. López, J A. Sanchidrián, A P. Santos

The cylinder test allows us to determine the expansion of a copper cylinder by the detonation of an explosive charge inside. This test, originally designed for military explosives, has been adapted for industrial explosives with non-ideal behaviour. From data obtained with this test and known physical relationships, an analytical procedure has been developed to calculate the Jones-Wilkins-Lee (JWL) equation of state parameters for five ammonium nitrate/fuel oil (ANFO) and six emulsion-type explosives. A new radial wall expansion function is proposed considering a non-zero initial velocity with the consequent positive Gurney energy at unit relative volume. A set of equations is established to determine the JWL parameters, three of them obtained at the Chapman-Jouguet (CJ) point and the other three at different relative volumes. These equations, along with some constraints to ensure a physically sound solution, are solved for the JWL parameters by a non-linear least squares scheme. A 2D Lagrangian axisymmetric finite elements (FE) numerical model is developed for each tested explosive to validate the analytical results, in terms of expansion histories, with good agreement. This confirms that the set of conditions, constraints on the parameters and least squares calculation used constitute a robust methodology that fully captures the physics of the detonation products expansion involved in the cylinder test.CITATION:Castedo, R, Sanchidrián, J A, López, L M, Segarra, P and Santos, A P, 2015. Jones-Wilkins-Lee parameters for civil explosives, in Proceedings 11th International Symposium on Rock Fragmentation by Blasting, pp 57–64 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 24, 2015