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By P Benjamin, H Bohannan, C Farr

Geoffrey Blainey's historical account of the mining activity on the west coast of Tasmania, æThe Peaks of LyellÆ documents the changes over 130 years of activity in the region. Commissioned by the Mount Lyell Mining and Railway Company, emphasis is placed upon the development of the Mount Lyell Mine in Queenstown. Chapter 19 æWidening HorizonsÆ details the review the Mount Lyell Company undertook in 1904/05 to ensure the future of the operation, following the pessimistic outlook regarding diminishing reserves and poor grade ore. The analogies documented between the Mount Lyell Company at that time, and Gold Mines of Australia 90 years on are remarkable, with both companies demonstrating a commitment to the Mount Lyell area through active exploration, the proactive search for alternative technology and the recognition of the valuable human resource. As Blainey writes (Blainey, 1993), æwith its expanding chemical works, its huge Tasmanian mines, its railways and smelter, its coke works and its mine crusade, Mount Lyell was probably one of the nation's five great industrial companies in 1906 or 1907. This brief phase of industrial and mining expansion had its seed in the threat of extinction . . .Æ After a number of reprieves, the Mount Lyell mine did stop production in 1994, but that only provided an impetus for a review and rebirth.

Jan 1, 1998

By P J. Harris, D J. Bradshaw

Surface oxidation of sulfide minerals, such as that found in the regions of a sulfide ore body near the water table, can have a significant impact upon flotation. This theme has been explored for Merensky ore type sulfides where an ore containing pyrrhotite, pentlandite and chalcopyrite was thermally oxidised and the role of potential remedies investigated. Back-scattered scanning electron microscope images are presented showing the oxidation layer which formed in the mineral surfaces. These oxidation layers were depleted in both sulfur and iron with incorporated oxygen. Flotation recovery rapidly decreased with increasing oxidation, particularly after 27 days and reached a plateau after 50 days. Up to 27 days, this effect could be partially overcome with higher collector additions. Oxidation had more impact upon the finer size fractions, particularly for pyrrhotite. For more heavily surface oxidised samples, ultrasonic treatment prior to collector conditioning was found to improve flotation recoveries. This treatment had the greatest effect upon chalcopyrite particles. Sulfidisation was successful in restoring the flotation recovery of the heavily oxidised sulfide minerals. Longer sulfidisation conditioning times were not conducive to good flotation recoveries of both oxidised pyrrhotite and pentlandite due to oxidation of the freshly formed sulfide surfaces. For maximum flotation recoveries of oxidised pyrrhotite, pentlandite and chalcopyrite, different sulfidisation conditions are indicated. It appears likely that in a mineral processing operation treating oxidised Merensky type ores, two stages of sulfidisation employing different conditions would be required.

Jan 1, 2005

The need to solve the problems associated with spontaneous combustion, system of ventil- ation and methane drainage in gassy coal mines has brought about the development of mining and degassing methods which has made possible up to 80% gas recovery in retreating longwalls. Beneficial results also were obtained mostly in the double or multiple-entry systems, but for single-entry systems and for a seam thickness exceeding 3 m, gas recovery rate is still very low. With some single-heading systems, such as those where a return roadway is not maintained in the goaf behind the face line, a new methane drainage technology is practiced. Attempts were made to extract methane from roof holes drilled from offsets in the ribside of the longwall block. The inclinations and lengths of the boreholes were planned and designed on the basis of the geological and mining condit- ions encountered at the time and every second hole was terminated precisely at the point where the previous hole intersected the immed- iate roof goaf area above the face. There are mathematical relations approp- riate to the design of the above drainage systems to give optimum methane recovery under the given conditions. Some examples of drainage have reached 87% recovery and have attained 94% CH4 purity.

Jan 1, 1982

By J J. Laas, A J. S Spearing

Considering the capital outlay, running costs and inherent difficulties of backfilling operations at depth the benefits need to be clearly established, quantified and then exploited. Expansion of Vaal Reefs Exploration and Mining Company towards even deeper, highly faulted ground, and the process of establishing codes of practice for the management of rockfalls and rockbursts, required that a sound support strategy be formulated. This paper describes the derivation of an overall backfilling strategy on Vaal Reefs. Use was made of experience gained over some 12 years with backfilling. Numerical modelling was used to estimate potential benefits of backfilling the deepest mining areas still to be extracted. In exploiting a probably unique opportunity, two contiguous mining areas bounded by the same major geological structures, were compared. One was extensively backfilled and the other not at all. Shaft sinking through a backfilled stope added another dimension to the database.

Jan 1, 1998

By I Manege, S M. Bennett

The PT Kaltim Prima Coal (KPC) operation is a large scale coalmine operating on the east coast of Kalimantan (Borneo), Indonesia. Originally developed under a coal contract of work (CCOW) with the Indonesian Government, new legislation has led to uncertainty as to the actual life of the mine, the size of the operating lease or mining area, the rehabilitation compliance requirements and other areas of concern. The actual life-of-mine (LOM) plan is therefore in a constant state of flux and must be updated and modified with every new piece of information.Under the CCOW system, the mine life ran from the first year of production (1992) for 30 years (2021), with the potential to extend for a further two ten-year periods. KPC has always intended to run until at least 2041. Under the new legislation, the potential term of the CCOW (to be replaced by a mining license) is open to interpretation, resulting in KPC having to prepare plans allowing for a LOM that may end in 2021, 2031, 2041 or a range of other possible periods of time.Uncertainty about legislation (and hence the mine life) possible production rates and available work areas; variable coal prices and costs; varying management contexts; the substantial reserves available; difficulty sourcing dump space; potential rehabilitation costs and the dependence of a community of over 100 000 people on a single operation sees the LOM and mine closure teams preparing dynamic yet thoroughly robust plans that are updated on multiple occasions per year.This paper will describe the procedure followed by KPC and the potentially show stopping issues that must be dealt with.CITATION:Bennett, S M and Manege, I, ?2014. Life-of-mine planning in a dynamic environment, in Proceedings Life-of-Mine 2014 , pp 379–388 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 16, 2014

Mining is still a significant economic activity within the Hauraki District. The Martha project is particularly significant. Estimates are that in the order of 850 direct and indirect jobs are derived as a consequence of the Martha project within the Waikato Region. The Resource Management Act 1991 (RMA) is New Zealand's primary piece of environmental legislation. It sets out to provide the framework for the control of the use, development and protection of almost all physical and natural resources. Minerals are excluded from the sustainability requirements of the legislation but as mining invariably involves the use of land, all the other elements of RMA apply. The Council's role in dealing with mining changed with the introduction of RMA. It became responsible for administration of those matters that fell within the functions set out in section 31 of the RMA in respect to existing authorisations. Mineral development needed to be considered in the development of the District Plan that the Council is required to prepare. Implementation of the Act has been a challenge for not only local government but for many industries and resource users since it came into force in 1991. There have been, and are, differing views on both the way it is administered and its intent. The Hauraki District Council has been at the forefront of some of the challenges that are a consequence of the Act. Its reactions to these challenges have been consistent. It treats mining in the same why it treats any other industry. At times this approach has not been understood and without challenge from sections of the community.

Jan 1, 2000

By R C. A Minnitt

The two methods for calibrating the parameters K and (alpha) for use in Gy’s equation for the fundamental sampling error (FSE), the duplicate sampling analysis (DSA) and the segregation fractionalised analysis (SFA) are described in detail. A case study using identical broken reef material from a Witwatersrand-type orebody is calibrated using the DSA and SFA methods and the results are compared. Classically, the form of Gy’s equation for the FSE raises the nominal size of fragments given by dN to a power of 3. Gy’s formula is a theoretical formula applicable in the case when fragments are randomly collected one by one, which is obviously never the case. This is because the fragment shape factor (f) and the granulometric factor (g) were introduced to take care of this anomaly. More recent formulae, such as those suggested by François-Bongarçon and Lyman, are empirically based formula. A later modification of Gy’s formula raises dN to the power ? (alpha), the latter term being calibrated with the coefficient K in the DSA and SFA methods. The preferred value of ? (alpha) for low-grade gold ores used by sampling practitioners in the mining industry is 1.5. A review of calibration experiments for low-grade gold ores using the DSA and SFA methods has produced values of K that vary between 70 and 170 and values of alpha in the range 0.97 to 1.30. The average value for alpha is found to be 1, rather than 3 as originally proposed in the classic form of Gy’s equation or the industry preference for 1.5. It is suggested that for low-grade gold bearing ores, the equation for the FSE should raise dN to a power of 1. Such an equation for the variance of the combined FSE plus the grouping and segregation error, now referred to as the short-term component of the quality fluctuation error (QFE1), greatly simplifies the characterisation of gold ores, which now only requires the calibration of K for a given mass and established fragment size. The implications for the heterogeneity test from the generalised equation are that provided the fragments have been screened to within a narrow size range, any particular size will return a value for K that is acceptable for use in the sampling nomogram.CITATION:Minnitt, R C A, 2017. A generalised form of Gy’s equation for gold ores – empirical evidence, in Proceedings Eighth World Conference on Sampling and Blending , pp 331–350 (The Australasian Institute of Mining and Metallurgy: Melbourne).

May 9, 2017

By R Minnitt

Two methods of calibrating the parameters K and Alpha for use in Gy’s formula for the fundamental sampling error were investigated. Approximately 480 kg of broken ore from Harmony’s Target gold mine in South Africa was used to compile the replicate samples for two experiments. The duplicate series analysis (DSA) and segregation free analysis (SFA) methods differ in terms of their approach to recovering the sized materials required for the calibration procedure. In this experiment the DSA method split off six more-or-less equal masses of material from a 75 kg lot that was successively crushed to finer and finer fragment sizes. The sublots so formed were split into 32 samples and fire assayed for gold. The SFA method used approximately 250 kg of broken ore crushed to 95 per cent passing 2.50 cm, which was then passed through successively finer screens until a sublot of less than 0.0075 cm was obtained. Each of the 15 sublots obtained in this way were split using a riffle splitter into 32 samples that were fire assayed. A total of 672 fire assays were used in these experiments. The sampling parameters K and Alpha were used for the compilation of a nomogram and to calculate the gold liberation size of the ores. While there are no definitive results in terms of the actual gold grain sizes against which the calculated sizes can be compared, it appears that the DSA method produces more acceptable results.CITATION:Minnitt, R, 2014. A comparison of two methods for calculating the constants K and alpha in a broken ore, in Proceedings Sampling 2014 , pp 165–178 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 29, 2014

By Jankowski P

The Bannockburn gold deposit is located 67 km north-west of Leonora in a poorly exposed part of the Norseman-Wiluna belt of Western Australia. The deposit consists of narrow, complexly shaped orebodies which have presented a challenge to find successful grade control and mining methods. The deposit also presents an opportunity to examine in detail the geology of this part of the Norseman-Wiluna belt.This paper discusses the geology and structure of the Bannockburn gold deposit as well as the grade control and mining systems developed to define the ore and extract it with minimum dilution. We discuss the different techniques employed for grade control in the past and their shortcomings, and the evolution of the present system of doing grade control in cross section from blast hole assays.

Jan 1, 1993

By A E Reschke

The high-shear colloidal mixer is generally recognised as the most efficient method of mixing cement based grouts. Only recently however has this technology been applied to the production of cemented rockfill (CRF) for use in underground mines in North America. Traditionally, paddle type mixers have been used to produce the cement based slurries required for the preparation of CRF. While considerable success has been achieved with these mixers, tremendous advantages are attained with the colloidal mixer. The combined effect of the highly efficient mixing action and the ability to mix low water/solids ratios, allows for reductions in the cement content for a given strength requirement. Cement can also be replaced by cheaper fillers such as flyash, resulting in further cost savings. Although large-scale dedicated CRF hatching plants are typical, there is a growing need for mobile plants capable of operating anywhere underground. The compact size and rapid hatching speed of high-shear colloidal mixers has allowed them to be incorporated into skid mounted systems capable of producing and discharging cement based slurries onto rockfill aggregate for the production of CRF. In many cases run-of-mine development waste can be used as CRF aggregate with these slurry plants thereby reducing the cost of waste rock removal. These relatively low cost portable plants have now allowed several mines to economically rccmer isolated pillars and realise positive returns on their investment.

Jan 1, 1998

Coal washery wastes comprise tailings, finer than 0.5 to 0.1 mm depending on the sophistication of the washery, and coarse reject, typically finer than 50 mm. Subaerial disposal of coal tailings slurry at about 30 per cent solids by mass results in the formation of a delta with an average slope of about 1 in 100, representing one extreme. The ravelling of coarse reject on loose-dumping, results in an angle of repose slope with an angle of about 37o (1 in 1.3), representing the other extreme. The pumped co-disposal of coal tailings and coarse reject results in the formation of an upper delta with an average slope of about to one in ten, between the two extremes. The depositional behaviour of these three waste streams is discussed and compared in this paper. The important depositional and material parameters, and a theoretical description of the processes involved, are also presented. Comparisons are made between the depositional behaviour of the various coal washery waste streams, and natural depositional systems, such as rivers, alluvial fans, and scree slopes.

Jan 1, 1998

By B Belle

In a deep, gassy predrained coalmine, ventilation and goaf drainage systems act as critical controls to manage longwall (LW) panel return gas levels. The goaf capture efficiency of most LW panels ranges from 45 to 85 per cent based on the gas domain at respective workings. Recent operational experiences have indicated that 50 m goaf hole spacing is optimal for management of high production retreat gassy LW panels, where the probability of goaf hole failures are frequent. Due to the horizontal stress direction geotechnical issues, goaf hole failure just behind the LW face is a recurring challenge resulting in significant and immediate increases in LW panel return gas levels. This paper presents a successful implementation of an innovative LW tailgate (TG) road cut through seal goaf fan/venturi system. The system incorporated a 17' (400 mm cased) hole drilled to 320 m deep in a perimeter roadway on a retreating LW TG road drawing the working seam horizon goaf gas behind the perimeter of seals. The surface goaf fan/venturi system was aimed to safely exhaust the LW deep goaf gas and thus withhold the goaf stream spillage to TG drive area and reduce the panel gas load by controlled operation of the system using continuous monitoring of CH4, CO and O2 levels. This control philosophy was made operational without impacting the use of the perimeter inbye ventilation shaft used for mine cooling and providing additional ventilation dilution capacity. The instantaneous reduction in TG gas levels of up to 0.3 per cent was observed even when the LW was 300 m away from the c/t seal goaf hole. Success of this concept TG seal goaf gas control system used even when the LW was 1300 m outbye has become an inventory of additional control for future deep and long LW blocks. This paper attempts to share the design process in addressing other risks and desired outcomes in managing the LW gas levels.CITATION:Belle, B, 2015. Innovative tailgate mobile goaf gas management in two gate road longwall panels – concept to implementation, in Proceedings The Australian Mine Ventilation Conference, pp 31–40 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 31, 2015

By P Bartlett, K Naidoo

Premier Diamond Mine is planning a new, mechanised, block cave, with a production level approximately 1000 m below surface. A high degree of tunnel damage is expected, due to stress and the relatively low in situ strength of kimberlite. An evaluation of the timing of tunnel excavation, tunnel size and spacing has been carried out to define critical parameters for planning layouts and development schedules. Extensive use has been made of three-dimensional FLAC3D numerical models, calibrated against known rock mass behaviour at 650 m depth, to examine the influence of undercut advance on production level excavations. These were able to represent the effect of stress rotation on tunnels during undercut advance and the effect of an inclined undercut. The analysis has provided guidelines that are possibly applicable to any deep block cave layout. The choice of drawpoint layout is of considerably less significance than the proportion of ground extracted on the production level elevation prior to advancing the undercut. The analysis shows that damage to production development will be potentially severe if it is all excavated prior to advancing the undercut, or is mined directly beneath the undercut face. The development of the main production tunnels alone ahead of the undercut may be feasible, but no breakaways should be put in as they appear to trigger a rapid increase in the severity of damage. The optimal sequence is when all production tunnels are developed at least 10 m in plan behind the undercut, where the induced peak field stress is less than 20 MPa and a maximum support pressure of 100 to 500 kN/m2 is require to contain movement. The assessment showed that the conditions are not significantly improved by increasing drawpoint spacing from 15 m to 18 m. The vertical spacing between the undercut and production level beneath is investigated. Provided production level development is carried out beneath the undercut, there would appear to be little benefit in increasing the vertical spacing between the undercut and production level beyond the current 15 m to 18 m.

Jan 1, 2000

By D J. Williams

Tailings management is typically constrained by many factors that are taken as given. These include the tailings production rate that must be accommodated, the desire to deliver the tailings cheaply and easily as a slurry using centrifugal pumps, and a tailings footprint limited by the desire to minimise capital costs leading to a rapid tailings rate of rise and a low settled density. Other constraints include wall raises being delayed due to limited capital, inadequate tailings water and rainfall run-off management, and the net present value (NPV) approach to capital expenditure that puts off expenditure at the expense of performance. All of these constraints result in the tailings occupying much more volume than is necessary, being difficult to rehabilitate, and not being capable of supporting a beneficial post-mining land use or function. Ultimately, such constrained tailings storage facilities will assume very large dimensions, but each will typically be developed as a series of smaller storages that are filled rapidly and added to progressively. The optimal stage to efficiently recover process water is in the process plant, but this is limited by the desire to use centrifugal pumps rather than the more costly positive displacement pumps, which also require greater management of tailings deposition. Process water can also be recovered from the tailings storage facility (TSF) decant. Tailings water that is not recovered from the plant or the decant will be lost to entrainment within the tailings, seepage to the foundation or through the wall, or evaporation. However, tailings water is not always suitable for reuse in the plant, particularly when it becomes contaminated by the processing, and it may simply be stored on the TSF. This will exacerbate the seepage of contaminated water to the environment, and will limit the densification of the tailings that would otherwise occur due to desiccation of the tailings on exposure. If a larger tailings footprint were developed early, cycling of tailings deposition in thin lifts between a number of cells could occur, maintaining the dormant cells dry to enhance desiccation and densification, and ultimate rehabilitation. The end result would be a much reduced stored volume of tailings, the potential for upstream raising, and ease of rehabilitation to a high level. The higher initial capital cost would be overtaken by reduced ongoing capital, operating and closure costs.CITATION:Williams, D J, ?2014. An alternative whole-of-life approach to tailings management, in Proceedings Life-of-Mine 2014 , pp 285–298 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 16, 2014

By S Gezgin, F Beckstein, M Müller, D Rapp, D Sergeev, J Qi, A Lichtinger

Salts and their mixtures play an important role for industrial and energy sectors, eg metallurgy, biomass gasification and combustion, nuclear and solar powerplants and electrochemical processes. Depending on the composition of salts (eg Li+, Na+, K+, Mg2+, Ca2+ // NO3-, F-, Cl-, CO32-, SO42-, etc), the temperature range for different applications can vary starting from room temperature and going up to 1500°C. For the proper design and modelling of heat exchangers and other equipment, it is necessary to have reliable and validated data sets of thermophysical properties. The 50 mol per cent NaNO3 – 50 mol per cent KNO3 salt mixture as a well-studied composition was selected for validation of relevant thermophysical properties (heat capacity, enthalpy of phase transitions, thermal expansion, viscosity and thermal conductivity). To study these properties of the liquid phase, special crucibles and approaches should be implemented. Verification of these crucibles for different methods (differential scanning calorimetry (DSC), thermomechanical analysis (TMA), laser flash analysis (LFA) and rheometers) has been performed. The low thermal conductivity of salts is one of the main problems in implementing latent heat storage. In this project, the authors intend to develop a fundamental solution to this problem using chemically bonded metals with salts. These are known as metal-salt solutions. To fit the main parameters of new phase change materials to the requirements of thermal energy storage, a wide variety of possible combinations should be considered. CALPHAD modelling is used together with thermal analysis for the development of a consistent thermodynamic database including chloride salts (Li, Na, K, Mg, Ca // Cl) and corresponding metals (Li, Na, K, Mg, Ca). The results of the study of quasibinary and multicomponent metal-salt (eg Mg-KCl and Ca-KCl) systems as well as the selection of suitable crucible materials and challenges by studying of these systems will be discussed.

Aug 21, 2024

In October 1993 it was decided to undertake the fourth phase of mining at Big Bell. This decision followed the exhaustion of open pit reserves in January 1993. Two stages of mining were proposed for the new underground mine. The initial stage would mine the remnant ore, consisting of the rib and crown pillars and lower grade ore on the hangingwall and footwall side of the old open stopes, which were mined from the 1930s to 1950s. Most of the old open stopes were backfilled with either deslimed mill tailings or development mullock, both of which were economic to recover. However, the mineable ore from this resource was difficult to evaluate. The second stage focused on mining ore outside the remnant areas and was initially proposed as a Core and Shell method under a caving hangingwall. This method has since been superseded with sublevel caving, in part due to the very high and deviatoric stress conditions encountered at Big Bell and the relatively long mining cycle required for open stoping. The problems and challenges discussed in this paper will cover the identification of remnant ore reserves, dewatering the old workings, and remnant mining.

Jan 1, 1998

The Emmaville district is located in the northern part of the New England region of New South Wales some 400 miles north of Sypney. Better known for its substantial production of tin and tungsten ores, the Emmaville district has also been the centre of a small and somewhat sporadic production of copper, arsenic, silver, lead and zinc. This production has come principally from three mines-Webb's Silver Mine, Webb's Consols and, to a lesser extent, Pye's Creek Silver Mine.The ores from these mixed sulphide deposits are texturally and mineralogically complex and the problems of beneficiation are rather formidable.GENERAL GEOLOGYThe main rock type of the Emmanville district is a dark, brown to blackish massive siltstone of Permian age which overlies acid volcanics possibly also of Permian age. In places the siltstone gives way to conglomerates and standstones with discontinuous lenses of calcareous sediments (David, 1887).The whole series is folded but because of the massive character of the more expansive siltstone, detailed structure is difficult to delineate.Toward the close of Permian time the area was intruded by a series of intermediate to acid plutonic and hypabyssal rocks. Mineralization within the region is ascribed to the more acid of these intrusives-granodiorite, granite, and quartz-felspar porphyry (Lawrence, 1957, 1960).No intrusive rocks outcrop in the immediate neighbourhood of Webb's Silver Mine though granite occurs two miles...

Jan 1, 1961

The cloud transform simulation is a technique used to generate realisations of correlated attributes. The principle of this technique is to partition the scatter diagram between the input attributes into several interval classes and derive for each class a cumulative distribution function (CDF) using only the data pairs of the scatter diagram in that class. The simulation then proceeds like a traditional Monte Carlo simulation but using the derived CDF in each class.The cloud transform simulation has the advantage of being a simple and fast technique that is able to reproduce the uncertainty in the relationship between the two input attributes regardless of whether the relationship is linear or not. However, the technique has three major drawbacks. First, the number of classes has an impact on final simulated results and there is not a robust approach to define such number. Second, simulated results exhibit some saw tooth variations that are unrealistic and third the reproduction of the spatial continuity of the simulated attribute is not guaranteed.This paper presents the use of the quantile regression approach in order to overcome the first two disadvantages of the cloud transform simulation. Details of the technique and how it can be coupled with the cloud transform simulation are presented along with a real case study.

Sep 26, 2011

Regional problems, such as political risk and debt reschedulings, in Pacific Rim nations coupled with the price volatility of gold mean that multi-faceted financing packages are needed to offset the many risks inherent in a new gold mine development. The cross-border risks of getting money back out of a producing mine may require the involvement of export credit agencies, multilateral government banks, or political risk insurance. Variations on forward gold sales, gold loans, or offset structures can also enhance the flexibility and safety for the providers of both the debt and equity. The region will become as large a gold producer as South Africa by 1990. The large dollar amounts for both investment and revenues realised are of great significance to the economies of many of the Pacific Rim countries especially as a means to increase foreign exchange earnings.

Jan 1, 1987

By D J Williams

Coal washery wastes comprise tailings, finer than 0.5 to 0.1 inn depending on the sophistication of the washery, and coarse reject, typically finer than 50 mm. Subaerial disposal of coal tailings slurry at about 30 per cent solids by mass results in the formation of a delta with an average slope of about 1 in 100, representing one extreme. The ravelling of coarse reject on loose-dumping, results in an angle of repose slope with an angle of about 37" (1 in 1.3), representing the other extreme. The pumped co-disposal of coal tailings and coarse reject results in the formation of an upper delta with an average slope of about to one in ten, between the two extremes. The depositional behaviour of these three waste streams is discussed and compared in this paper. The important depositional and material parameters, and a theoretical description of the processes involved, are also presented. Comparisons are made between the depositional behaviour of the various coal washery waste streams, and natural depositional systems, such as rivers, alluvial fans, and scree slopes.

Jan 1, 1998