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By C E. Loo

The cohesive zone has a significant influence on blast furnace operation and performance. The high-temperature behaviour of the reducing ferrous burden determines the properties of the cohesive zone and is studied in a laboratory using a softening and melting test. This paper summarises recent work aimed at better understanding the physico-chemical changes in a ferrous bed during softening and melting. Using a combination of standard, non-standard and quenched tests supported by optical image analysis, this work has highlighted some important differences in behaviour between lump, pellets and sinter as well as between the two lump ores and two acid pellets. As with the two lump ores, one pellet is denser than the other and the denser ore/pellet is also higher in silica.Within the cohesive zone, densification of the ferrous layer leads to reduced permeability and greater pressure drop. From the work presented in this paper, two key mechanisms of ferrous layer densification were identified. Firstly, the formation of fayalite, which melts at low temperatures, can act as a lubricant between particles and lead to densification of the ferrous layer under load. Secondly, reduction to metallic iron is topochemical and fayalite in the outer shell of the particle restricts gas flow to the centre and results in the formation of a wustite core. As temperatures rise, the wustite melts and eventually exudes out of the shell, occupying void spaces between particles and further reducing layer voidage and permeability.This work has shown that density and silica content are two key factors that result in increased densification of the ferrous layer. For porous Australian lump ore, the higher porosity and lower silica reduces the formation of fayalite and limits the formation of the wustite core, ultimately leading to a more permeable bed. The difference between the porous and dense ores was more pronounced with the introduction of prereduction to better simulate blast furnace conditions. This suggests that a porous ore gives significantly better cohesive zone permeability and blast furnace operations than a dense ore.The behaviour of fluxed sinter particles is different in that the silica reacts with lime to form dicalcium silicate. Minimal wustite core forms because relict hematite particles in sinter are only small and reduction occurs uniformly throughout the particle. Even with significant loss of material through prereduction, beds formed from sinter remained comparatively rigid. Studies using a typical blast furnace burden mix of 80 per cent fluxed sinter and 20 per cent porous ore showed that the structures of these beds at softening and melting temperatures are very similar to the sinter-only test. Measurements of bed contraction and pressure drop also confirmed that the introduction of 20 per cent porous ore into a sinter bed did not alter its properties. This is an indication that 20 per cent ore can be easily accommodated by blast furnace burdens.CITATION:Loo, C E, 2015. Important acid ferrous burden properties in the cohesive zone of a blast furnace, in Proceedings Iron Ore 2015, pp 367–376 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 13, 2015

EQUIPMENT TO BE DISCUSSED It is the intention of this paper to discuss the transporting of large skid mounted equipment and track machines around open cut mines. By large I mean 100 to 350 tonne capacity being trans- ported on minesite lowloaders and the capability of moving loads of 300 to 1200 tonne by other means. It is also possible to move larger loads. The main factor in transporting large loads is eco- nomics and is a balance between the cost of the transporter and the increased production from the equipment being moved. For example it could be possible to shift a 7000 tonne dragline faster than their own means of propulsion, being the "walking feet" method. However for the cost of $10 to $12,000,000.00 or more you can afford to walk the dragline and suffer the production loss. Let us concentrate on the economically realistic. BOSICH FBT MINESITE LOWLOADERS 100 TO 350 TONNE CAPACITY. Why initially consider having such a transporter? Our experience has shown that initial enquiries evolve from mines requiring to move large drills, 50 to 200 tonne weight and large hydraulic excavators or backhoes which are track mounted. The reason for shifting them is to move equipment from a mining face to enable blasting, to remove to the central workshop for repairs, overhauls and routine maintenance, to transport equip- ment from one mine face to another, and for drills, to move from one drilling pattern to the next. The cost in terms of time lost and extra wear on tracked equip- ment easily justifies the cost of the transporter. The tracks on large drills and shovels are not intended for long distance walk- ing. These tracks are designed to move the machines to position for drill holes, or moving around the mine face. Long distance travel, ie generally 1 km or more, results in overheating of track rollers, track drive components and accelerated wear on tracks.

Jan 1, 1989

The Mount Kasi gold deposit in Vanua Levu represents an epithermal, high sulphur, gold-enargite system occurring within a volcanic caldera of Middle to Upper Miocene age. High level breccia pipes flare upward and form flat lying, near surface, mushroom shaped sheets parallel and near to the present land surface. The upper portions of the pipes contain explosion breccias which are capped by sandy siliceous sinter-like beds and are cut by alunite veins. Outlying intercalated scoriasceous lavas and breccia sheets show evidence of subsidence within a shallow basin. Detailed mapping of breccia types, intensity of silicification and the presence of multiple quartz veining and enargite abundance outlined blocks of more consistent gold mineralization within the breccias. This work has provided a model for mineralization which will aid future exploration and ore reserve estimations. A significant eluvial concentration of gold occurs in soils overlying the breccia sheet.

Jan 1, 1987

By Roberts C

The satisfactory testing and characterisation of mineralisation amenable to heap leaching, is of prime importance to successful operation of a project. This paper outlines a methodology for assessing material suitability. A general model is used for classing material types for suitability for treatment and for economic analysis. The four levels of these are overburden, oxide, transitional and unoxidised materials. During exploration, reverse circulation (RC) drilling can be used effectively to define these zones by a combination of logging, analysis and provision of cuttings for column leach and bottle roll trials. RC cuttings can he used effectively for column leach trials, when calibrated against conventional processing-sized samples, obtained from surface. Tbis allows metallurgical testing to depth and definition of transitional ore types and the base of oxidation, accurately and economically with an acceptable density of samples. The accurate prediction of base of oxidation and variations in recovery with depth and ore types can be used to develop 'economic' pit designs with zones of variable recovery. Petrological studies of leaching results and examination of mineralisation types carried out on concentrates, provide information on gold distribution and physical and chemical constraints. The examination of concentrates can be particularly useful for types of mineralisation when overall gold content is frequently at or below one pan per million. Explanation of anomalous recovery of gold grains in dense goethite and solution-robbing by carbonaceous material can be identified by use of petrological and scanning electron microscope. An understanding of the role of the host rock in leaching characteristics can also he provided by petrology where siliceous or clayey materials can require significantly different treatment.

Jan 1, 1994

By Wright J. K

In recent years the production of steel from directly reduced iron ore pellets has ex- panded rapidly. However continued growth will depend on an adequate supply of suitable raw materials for the various commercial gaseous and solid reductant processes now in operation. Al- though at present Australia does not produce pellets primarily for direct reduction, there is an increasing awareness of Australia's potential role as a supplier of high grade ores for export and/or domestic use. The properties and quality requirements of iron ore pellets made specifically for directre- duction (DR) are reviewed and the laboratory techniques and test methods used at the CSIRO Division of Process Technology for the evaluat- ion of potential DR feedstocks are described. At the present time most of the evaluation tests for DR pellets are extensions of those de- signed for existing blast furnace (BF) burdens. In the case of some physical parameters such as cold compression strength, abrasion index and cold strength after reduction this extension is quite satisfactory. In the measurement of para- meters such as reducibility, swelling and stick- ing tendency, however, it is necessary to deve- lop meaningful standardized tests so that ores and pellets can be ranked for DR performance. To do this, data from the laboratory tests and from operating plants must be correlated.

Jan 1, 1980

By K Dagdelen, I Traore

This paper presents an iterative net present value (NPV) maximisation method to determine the optimum surface to underground transition depth for an orebody to be mined by multiple open pits and an underground mine. The determination of transition depth from open pit to underground mining is based on global production scheduling optimisation of open pit and underground mines using mixed integer linear programming. The method is applied to a case study coming from a gold mining complex with six open pits and a large underground mine using longhole open stoping. The results indicate potential improvements of the NPV of global operations when compared to the traditional techniques based on independently optimised open pit first, followed by underground mining.CITATION:Dagdelen, K and Traore, I, 2014. Open pit transition depth determination through global analysis of open pit and underground mine production scheduling, in Proceedings Orebody Modelling and Strategic Mine Planning Symposium 2014, pp 195–200 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 24, 2014

By J E. Lynch

The Mt Leyshon gold deposit produced in excess of three million ounces of gold and considerable silver over a period of approximately 130 years. Most of this production occurred between 1987 and 2001. Mining commenced in 1872, however it was highly likely that Chinese miners were not only removing alluvial gold from the area but were also mining very rich æreefÆ gold. Records for early alluvial and hard rock mining do not exist, however gold production was probably considerable. In the early 1980s Marathon Petroleum (Australia) Ltd (MPAL) joint ventured into the Mt Leyshon area which had been held as a mining lease by Noranda Australia Ltd since 1968. Noranda at that time was concentrating its exploration effort on the Balcooma base metal prospect areas to the north of Charters Towers. In early 1982 MPALs parent company was taken over by US Steel. This led to an opportunity for MPALs Australian management to purchase all tenements and licences (except for coal, oil and gas) which were duly purchased for $US 1 million. At this time MPAL had a remarkable array of gold tenements (many of which were subsequently mined by other companies) and Mt Leyshon was considered to be the highlight. Early æseedÆ capital ($A 3.1 million) was raised with backers being the Bank of America, the NSW Government Superannuation Scheme, Westpac, private individuals and a Queensland grocery chain by the name of æJack the SlasherÆ.

Jan 1, 2004

By V Ser, I Smit, C Robertson

Anglo American has been researching the potential benefits of inter-particle comminution techniques for several years. Inter-particle comminution has additional benefits in terms of milling circuit performance such as energy consumption and stability. This paper focuses on the flotation benefits. Inter-particle comminution leads to selective breakage, mineralogical analysis has shown that this leads to improved liberation of the valuable minerals. In addition to selective breakage, the relatively short residence time experienced in typical inter-particle comminution machines allows more efficient classification which contributes to a æsharperÆ product size distribution. The effect of particle size in flotation has been well documented in literature. In general, it is has been shown that there is an optimal particle size range for flotation, and that both above and below this range, flotation recovery is adversely affected. Results show that the narrow size distribution which results from inter-particle comminution improves flotation recovery. Furthermore, the absence of excessive fine gangue material which typically results from conventional fine (-75 ¦m) milling of talcaceous ores, can lead to 100 per cent savings in depressant consumption when such ores are subjected to inter-particle comminution. To-date, this technology has been evaluated on more than ten ores. This paper will focus on one set of results which demonstrates a number of common differences that exist between material that has been milled conventionally and using inter-particle comminution technologies, with emphasis on what the impact of these differences is on flotation performance. A pilot scale (200 kg/h) vertical roller mill (VRM) was used to produce the material, and the results shown will describe conventional laboratory batch, locked cycle and pilot plant flotation testwork.

Jan 1, 2005

The project is undertaken by Elgas Pty Ltd, a 50 per cent joint venture between AGL and BOC, to store 65 000 t of LPG in underground storage at Port Botany, NSW. The contractor, Concrete Constructions Pty Ltd, elected to use drilling and blasting techniques. The project specification called for the following works to be excavated: Access Shaft:122 m (out of 140 m total depth), 6.0 m diameter (selected by the contractor),Operation Shaft:125 m (out of 150 m total depth), 4.0 m diameter,Water Curtain Gallery:500 m (Later extended by additional 270 m), 4.0 m wide and 3.5 m high,Storage Caverns:910 m, 14.0 m wide, 11.0 m high,Top Heading:14.0 m wide, 5.5 m high,Bench:14.0 m wide, 5.5 m high,Connection Galleries:250 m, 5.5 m wide, 5.5 m high.A total of 156 000 cubm of rock to be excavated.

Jan 1, 1999

By H Bartosiewi, M Kuruppu, F Mashiri

Wire ropes are one of the most heavily used components in a dragline. They are subjected to harsh conditions during the regular usage of a dragline in a mining operation. Hoist ropes are subjected to fatigue due to the cyclic nature of load handling as well as due to rope bending over the sheaves and the drum under load. This leads to wire breaks due to fatigue. Accumulation of a number of wire breaks close to each other can have a detrimental effect on the rope. Furthermore, to allow for the increasing demand for higher load capacity coupled with the inconvenience of having very large ropes, the factor of safety is often compromised, which increases the wear rate. Drag ropes are also subjected to heavy loads. More importantly, they are allowed to drag along the rough mine surface subjecting them to external physical abrasion. This makes the life of drag ropes one of the lowest among those used in a dragline. Suspension and IBS ropes are relatively uniformly loaded during their regular usage although they need to withstand dynamic load cycles as well as bending. Hence they tend to last for a number of years on average. The paper analyses the wear types and their severity of each of these rope applications, and suggests methods to determine rope wear rates and the resulting rope life. The paper further gives suggestions for good operating and maintenance practice that can extend the rope life and help reduce the large expenditure associated with every major rope change in a dragline.

Jan 1, 2005

Longwall mining has historically been considered an application for only deep reserves. This has changed in the recent past as longwall has become competitive with surface mining. Many resources marked for stripping now he in the domain of longwall. Today, the best longwall mines compete with strip mines operating in even the lowest stripping ratios. Longwall mining has become more widely applicable than it ever has been in the past and it is now generally considered the method of choice in situations where high levels of coal production are required. But the best way to use longwall mining has become less certain with the opportunities that highwall access provides. Wall-to-wall mining is the use of retreating longwall developed from and retreating toward a highwall. As a surface mine reaches its ultimate highwall, the opportunity for longwall punch mining or wall-to-wall mining is obvious. There are several clear operating advantages to operating a longwall in this way and where highwalls are available, the best economic decision will generally favor planning a direct approach. Yet wall-to-wall is not applicable in every situation even when highwalls are available. Moreover, the advantages of wall-to-wall mining do not always justify the cost of trenching or box cuts to access coal . However, many tactical advantages of operating from the highwall can be realized with the use of specially equipped blind drilled or raised bored shafts. This paper discusses novel techniques for a shaft assisted mining system which permits the addition of conveyor structure and mine systems from the surface to the face for advance and retreat mining with longwall, shortwall or room and pillar panels.

Jan 1, 1998

By P Ireland, S Mitra, G Evans, A Wang, M M. Hoque

This work presents an experimental study of flow dynamics in a bubble column system (crosssection: 100 mm × 100 mm) using a non-intrusive optical technique with targeted application in a novel REFLUX™ Flotation Cell. The study had two aims – measurement of the bubble size distribution (BSD) and local energy dissipation rate in the system due to bubble motions over a low range of gas flux varied from 0.02 to 0.08 cm/s without impairing the visibility of the system. A central bubble plume was generated by injecting compressed air through a sintered glass distributor in quiescent liquid. Highspeed shadowgraphy imaging was performed to determine the BSD in the presence of surfactant (Sodium Dodecyl Sulfate (SDS) concentration ~ 10 ppm). The BSD exhibited a unimodal shape at different locations of the column for all gas flux cases. A non-intrusive particle image velocimetry (PIV) technique was used to quantify the flow field generated by the bubble motions, which was used to estimate the local energy dissipation rate. The local energy dissipation rate exhibited an increasing power-law dependency on the gas flux.

Aug 24, 2022

By G J. Jameson, A Agarwal

This paper is concerned with the observation and analysis of bubble clusters in a laboratory flotation cell. A device was used in which bubbles were allowed to rise in clean water containing the same frother and collector concentration as in the cell. Bubble clusters were recorded using a high speed camera and analysed using an image analyser. Dodecyl amine was employed as a collector to float silica. The effect of impeller speed on the cluster behaviour was investigated. The results were interpreted in terms of cluster diameter and shape factor. The sizes of the bubbles forming the clusters were also measured and compared to the sizes of the independent bubbles in the cell. It was found that there was an optimum impeller speed in which the cluster diameter was a maximum. Higher agitation and the consequent higher power inputs resulted in cluster break-up while lower shear rates promoted cluster formation.

Jan 1, 2005

By S C. Bhowmik, R Gupta

This paper presents the ventilation problem faced by Jaduguda uranium mine, operated up to a depth of 905 m. The first two stages of the mine workings are up to 555 m, which are nearly exhausted. The third stage of the mine is from 555 m to 905 m and presently most of the stopping and development activities of the mine are carried out in the third stage. The total available intake quantity in the third stage was only 21 m3/s. As a result, the mine was facing a radiation hazard. The different instruments and the methods used for monitoring of radiation levels are also presented in the paper. It was observed that the average individual effective dose (external + internal) of radiation workers in Jaduguda mine increased from 6.93 mSv/y in 1995 to 9.5 mSv/y in 2001 and during this period the average individual effective dose of mine workers of the world was less than 5 mSv/y. Keeping this problem in view, a detailed ventilation investigation has been carried out in this mine. Initially the requirements of intake air quantity have been assessed by carrying out field investigation in different stopes and it was found that the third stage workings require nearly 100 m3/s. A detailed ventilation survey has been carried out in the total ventilation network and a basic ventilation model of the mine has been prepared. Different alternative schemes for effectively ventilating the third stage workings of the mine were simulated and a set of recommendations were made for implementation. The recommendations were partially implemented in 2002. As a result the total intake quantity to third stage workings increased to nearly 70 m3/s and the average individual effective dose (internal + external) of mine workers came down from 9.5 mSv/y in 2001 to 6.65 mSv/y and 6.60 mSv/y in 2002 and 2003 respectively. The implementation of remaining recommendations is in progress. Finally, a mathematical model has been formulated from the first principle for predicting air quantity requirements in cut and fill stopes in uranium mines and this model is being validated for further use.

Jan 1, 2005

By S Pigram, G J. Jameson

This paper is concerned with the tracking of particles in the froth phase of an aerated water/glycerol mixture. Experiments have been carried out in a specially designed laboratory flotation cell that allows formation of a deep froth. Phosphorescent tracer particles of various sizes were injected into the centre of a froth column where they were excited by ultraviolet lights, positioned at the back and front of the column. The motion of particles was captured on a digital camera with a green filter. The images from the digital camera were then transferred to a computer and an image analysis program was used to convert the colour intensity to the concentration of particles at each location within the froth. The tracer technique was used to determine the dispersion of hydrophilic particles and the variation of the concentration of solids with axial and radial positions.

Jan 1, 2005

Prior to 1975, most geochemical drainage surveys for gold were based on analyses for pathfinder elements, such as As and Cu, rather than direct analysis for gold, for two reasons. Firstly, analytical methods lacked the required sensitivity and precision to determine gold abundance at background levels in stream sediments (usually in the range 1 - 10 ppb). Direct analysis was also inhibited by concern about the potential for gross sampling error, brought about by the dispersion of gold as rare discrete particles of the extremely dense native metal. This sampling error, commonly termed the 'nugget effect' has two components: erratic distribution of the element within the stream environment, related to hydraulic factors, and the difficulty in taking representative samples, both in the field and later in sub-sampling for analysis. In more recent years the collection of very large samples was commonly advocated to overcome the sampling problem (Hams, 1982; Nichol, 1986).Two methods have been widely used to evaluate the gold content of bulk samples (Elliott and Towsey, 1989). The preparation and analysis of heavy mineral concentrates has in recent years been largely superseded by the leaching of gold contained in bulk samples by cyanide, commonly termed the BLEG (Bulk Leach Extractable Gold) or BCL (Bulk Cyanide Leach) methods. The terms BLEG and BCL are essentially interchangeable and the latter will be adopted for the remainder of this paper. The strengths and weaknesses of these two techniques were compared with conventional total gold analysis on fine fraction (

Jan 1, 1976

The geotechnical factors controlling the design of the open pit slopes of the Pasminco Century Mine near Lawn Hill, Queensland are presented below. Pasminco Century is a stratiform shale hosted zinc deposit in the Lawn Hill Formation that is overlain in part by the Thorntonia Limestone Formation. Both the mineralisation and overburden is located within a competent rock mass such that intact rock strength and discontinuity spacing is generally sufficient to prevent large-scale slope failure by breakage through the rock mass. Rather, overall slope stability issues are dominated by the major discontinuities within the rock mass. The deposit is located in a structurally complex regime, it consists of a synclinal basin which is truncated by easterly striking normal faults through its centre and northern and southern limits and by an unconformity on its eastern and western margins. Deposit geometry and slope stability is further complicated by small-scale folding, a pattern of northeast and southeast striking subvertical faults and north easterly dipping thrust faults with their associated fault emplaced blocks of Lawn Hill Formation in the overburden. The large size of the deposit requires a staged mining approach with the excavation of both final pit walls around the external sides of each stage and interim pit walls along the internal boundaries between adjacent stages. The deposit geometry and major discontinuities control not only the final pit wall design but also interim wall location and orientation, excavation sequence and in some stages the excavation schedule. Two and three-dimensional limit equilibrium stability analyses were conducted using both deterministic and probabilistic approaches to determine appropriate pit wall design criteria and to select a suitable excavation sequence and schedule.

Jan 1, 2000

By B Godel, E R. Ramanaidou, M Verrall

Fortescue Metals Group Ltd Iron Bridge (IB) operation is a magnetite deposit located in the Pilbara region of Western Australia situated around 145 km south of Port Hedland. This Mesoarchean deposit belongs to the 3230 Ma Pincunah banded-iron formations (BIF), a member of the Kangaroo Caves Formation, Sulfur Springs Group (van Kranendonk, 2000).

Sep 18, 2023

By P Bethune

Back Creek is a tributary of the Condamine River. It flows in a northerly direction across the existing Commodore coal reserves. The diversion of Back Creek was proposed to minimise sterilisation of the reserves. The approval process required a Waterworks Licence application, under the Water Resources Act 1989 (Qld), for the construction of a stream diversion channel and floodway. The diversion was designed using the æreference reachÆ approach, which mimicked existing properties of Back Creek. The following design standards were developed for the diversion: the diversion corridor width was 500 m; the diversion geometry was specified for both the low-flow channel, and the floodplain; material properties were specified for both the low-flow channel lining and the diversion bund; compaction standards for both the low-flow channel and floodplain were set; and the diversion rehabilitation included a riparian corridor and pasture land with clumps of native trees and shrubs. The following performance criteria were developed for the diversion: velocity regimes were replicated for the design events considered, in an attempt to avoid increasing the diversion flow velocities in comparison with the existing creek flow velocities; sediment transport calculations indicated that no significant change in sediment transport was expected within the diversion; overburden settlement estimates for the spoil placed in the corridor were summarised, and will be monitored throughout the construction process; geotechnical stability of the proposed Eastern Pit Buttress and the diversion bund configurations was confirmed; and groundwater impacts were assessed. Construction techniques, monitoring and further investigations are also discussed in this paper.

Jan 1, 2003

Mining is a high risk, high capital and potentially high reward undertaking. A large proportion of the risk comes from a lack of definition of the orebody (which also provides some of the potential) and the unpredictable nature of the market. This makes the planning process difficult but encourages the development of a robust yet flexible mining system. The system should be capable of absorbing some deviation from the anticipated concept without a significant impact on profitability. The mine planning process tends to consist of a series of related but discrete stages. For instance, the pit shells are designed, a schedule for the pit shells is created, a plant fleet is chosen to achieve the schedule and a financial analysis is carried out on that scenario. It is likely a few iterations are carried out using various mill throughputs to produce a favoured financial result. The problem is that handing results from stage æBÆ for use in æCÆ doesnÆt encourage a flexible system. æCÆ is likely to be built for exactly æBÆ. Unfortunately, in mining, æAÆ (the orebody) has the highest potential for variability and impacts everything else downstream. æDÆ the market also has high potential for unplanned variability. These variations can spread and æreverberateÆ through the whole mine process. For example an unexpected increase in ore oversize impacts all downstream mine processes to the primary crusher which may in turn create problems back æupstreamÆ by restricting throughput rate. Also an unexpected drop in demand or price could require modifications to the planned production rates. This paper looks at how mining process simulation can be used to support the development of strategies to counteract unforseen circumstances or alter production to match fluctuations in demand and price.

Jan 1, 2003