Search Documents

By Denholm W. T

The SMELT-2 package is specifically designed to perform on-line simulation of smelting processes. This paper describes its application to the simulation of continuous matte converting followed by reductive slag cleaning in SIROSMELT reactors. The matte, containing 50% Cu is assumed to be produced at the rate of 20 tph in a continuous matte smelting furnace such as an Outokumpu Flash Furnace, Noranda or Sirosmelt reactor. Two alternative fluxing strategies in the converter are examined, namely: ò conventional silicate slags, ò calcium ferrite slags, and the flow sheets associated with each strategy are discussed. A regular solution model is used in the present study to account for the variations in the activity coefficients of components in the slag.

Jan 1, 1989

The Que River and Hellyer polymetallic massive sulphide ore bodies are located 3 kilometres apart in the Cambrian Mount Read volcanics of Western Tasmania. Both deposits are classic examples of the Kuroko volcanogenic style. Base and precious metal rich hydrothermal solutions rose through a deep fracture system and vented on the sea floor, causing sulphides to accumu- late in adjacent topographic lows. Together the preproduction geological resource is approximately 23 million tonnes at greater than 18% zinc plus lead, with economically significant silver and gold. Exploration by Aberfoyle in the Mount Read volcanics com- menced in 1970. Que River was discovered in 1973 following a comprehensive data review, a regional stream geochemical survey and a helicopter-borne electromagnetic survey. Production began in 1981, and the mine is currently producing 300,000 tonnes per year. The ore is sold to the Electrolytic Zinc Company of Aus- tralasia Ltd mill at Rosebery. Ten years of extensive geological mapping, soil geochemistry, and induced polarisation, combined with continual study of the Que River orebody and its host rocks, formed a strong frame- work for the discovery of the Hellyer deposit, which was found as a result of drilling .a subtle time domain electromagnetic anomaly in 1983. Hellyer is currently being mined at a rate of 250,000 tonnes per year, with the ore being concentrated at Aber- foyle's converted Cleveland tin mill. Treatment of 1 million tonnes per year will begin in early 1989 at a new mill located near the Hellyer site.

Jan 1, 1988

By I Grewal, B Klein

The Knelson Continuous Variable Discharge (CVD) is a continuous centrifugal gravity concentrator for high-mass yield recovery applications. It has four operating variables that enable control of mass yield, product grade and recovery namely: fluidisation water, bowl speed, pinch valve open time and pinch valve closed time. Plant testing of a pilot scale Knelson CVD6 was conducted to evaluate the effects of operating variables on mineral separation. The analysis of operating variable effects has provided a framework for the development of operating guidelines for the CVD.

Jan 1, 2001

By P Rubie, J Russell

This paper looks at the past, present and not too distant future of grinding mill relining.Since the adoption of autogenous grinding (AG) mills and semi-autogenous grinding (SAG) mills as the primary ore reducing devices in the comminution circuit, plant shutdown frequency and duration has largely been driven by the liner maintenance requirements of these mills.The past 30 years have seen the introduction of many technologies to reduce relining times and frequencies. Advanced relining systems have been developed which include single and twin Mill Relining Machines, capable of placing large (up to 8000 kg) liners, high energy recoilless THUNDERBOLT Hammers, and high capacity Feed Chute Transporters. Despite these advances, today many of the hazards associated with mill relining remain and reline times still vary considerably between mine sites. A sophisticated discrete event simulation program, designed to analyse the mill relining process, has revealed the dominant influences on reline times of crew performance and of the need for safe access to larger segments of the mill.A new generation of advanced and semi-automated technologies is currently being tested that will address these remaining challenges. These technologies remove people from the hazardous environment inside the mill and will allow techniques used in modern advanced manufacturing to provide a near-future step change in mill relining safety and performance.CITATION:Rubie, P and Russell, J, 2018. 21st century relining – the path to greater mine site profitability, in Proceedings 14th AusIMM Mill Operators' Conference 2018, pp 621–640 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 29, 2018

By Kavetsy A

In surface mining operations, the ease with which a muckpile may be dug is a dominant factor in mining efficiency. Within broad limits this is largely independent of the fragment size dis- tribution, since the equipment used to move the broken rock is generally larger than in underground mines. In such cases, muck- pile shape and voids ratio within a muckpile heavily influence mine productivity. In a particular operation, the muckpile shape after blasting should be tailored to the digging equipment being used. A front- end loader is best suited to a low, flat-lying, loose muckpile, because of its fairly good mobility and small break-out force with its bucket. In contrast, a rope shovel lacks mobility, generates a much larger break-out force with its bucket, and fills its bucket in a vertical filling motion. Therefore, it is ideally suited to well fragmented, steep faces. In situations where a dragline is used, the requirement for a suitable muckpile depends on the operat- ing scheme of the dragline. For a conventional highwall key method the preferred work- ing level of the dragline is the same as the highwall, so that the dragline may sit over its key. In this situation, the ideal muck- pile shape involves relatively little displacement of the fragmented rock, with a minimal "power trough" at the back of the blast. In throw blasting, on the other hand, it is desirable to cast as much of the spoil as possible across the pit. It is then necessary to optimize the blast design by maximizing the throw percentage under any necessary constraints (e.g. damage, vibration, flyrock, explosive cost).

Jan 1, 1988

By J Arola, B Kettunen, W M. Bond

Separation of fine and coarse particles as separate streams prior to flotation, and their differential treatments, could substantially increase recovery and reduce collector consumption. A series of simulations was carried out to quantify the benefits from plant scale implementation of such a treatment at the Ispat Inland flotation circuit (Minorca mine). The study involved detailed plant sampling and sample analysis to generate the data needed for simulation studies. Samples were taken from all the streams from flotation feed to final concentrate, as well as cell lips and cell pulp. Sampling was repeated for the second ore blend that was processed by the plant during different periods of a year. Sample analysis involved sieve analysis and size by size chemical analysis for total iron, magnetic iron and silica. Raw data were mass balanced, and flow rates within the circuit were calculated. Mass balanced data were used to calculate model parameters representing the existing conditions. Usim Pac was used for both mass balancing and simulation studies. Available pilot and bench scale data were utilised to define the modified flow sheet and expected performance. The simulated flow sheet included classification of flotation circuit feed by hydrocyclones. While the coarse fraction followed the existing flow sheet, the fine fraction was sent to a magnetic separator. Magnetic and rougher flotation concentrates were combined to form the final concentrate. Two simulations for each blend were carried out. The first one assumed that there would be no change in size by size kinetic behaviours of both gangue and magnetite due to this modification. This was named as the worst case scenario, in which particles had a longer residence time due to reduced flow rates within the flotation cells. The second scenario anticipated that there would be an increase in recoveries of gangue and magnetite going into froth launder due to more efficient use of collector in absence of fines. Results showed that pre-classification of flotation feed could improve the flotation recovery by 0.3 - 0.5 per cent, with the major benefit being significant reduction (40 - 60 per cent) in the amount of collector used.

Jan 1, 2005

The gas drainage characteristics of two seams, the Bowen at Bowen No. 2 Colliery, Collinsville, and the Gemini at Leichhardt Colliery, Blackwater, are assessed. Single borehole tests and drainage panels were studied by underground monitoring of pressures and gas flows. Permeability was found to be effected by cleat to an order of magnitude. Other structural features are shown to be important. In this context, the need for delineation of no-flow zone locations in boreholes is shown to be of great importance where in-seam pre-drainage is being used for outburst control. Water in the seams strongly influences permeability causing gas flows to fluctuate with time. Use is made of modified reservoir engineering techniques to analyse the drainage characteristics of seams.

Jan 1, 1982

Alternative processes for converting coal to liquid products and their stage of development are outlined. The merits of hydrogenationprocesses in general, and the solvent refined coal process in particular, are stressed. Bench scale experimental hydrogenation procedures used by the Australian Coal Industry Research Laboratories Ltd. (ACIRL) to study the liquefaction of Australian coals and correlations between properties of such coals and their conversion to liquid products are described. Factors involved in assessing investment and operating costs for modern hydrogenation based coal refineries of different capacities are described with reference to a recent study undertaken in N.S.W. The significance of coal properties such as yield and yield structure of total liquid products and how they and other factors can influence the economics of the coal refining operations are discussed.

Jan 1, 1977

In recent years a large amount of work has been done at various plants overseas, and also at Broken Hill Associated Smelters, Port Pirie, South Australia, on the use of Dwight Lloyd sintering plant gases for sulphuric acid manufacture by the contact process.Several plants have modified their sintering machines and operating procedure, and some have adopted gas recirculation techniques, to obtain a relatively high proportion of the sintering plant gas at an S02 concentration acceptable to a contact acid plant 1.These adaptations were investigated at Port Pirie, and it was found that gases of suitable S02 concentration could be obtained. The results justified installation of a contact acid plant and a decision was accordingly made in 1949 to construct such a plant.However, the local conditions presented some difficulties not generally encountered at other plants, the chief differences being:a. For maintenance of production of 98% H2S04 , an unusually high S02 concentration is required, because of the prevailing high temperature 'of the purification train cooling water.b. The entire sintering operation is conducted on two large machines, 55ft. x 10ft., operating in series as "First" and "Second over" machines (a two-pass roast.) Any interruption" to either machine causes a complete interruption of plant operation and gas flow.

Jan 1, 1951

By Kwok S, Haemmerle A

Treatment methods for a low grade deposit at Mt. Fitch near Rum Jungle, N.T., were investigated. Physical beneficiation of the ore, sulphuric acid leaching, carbonate leaching and various methods of uranium extraction from the leach liquor were studied. Although the deposit occurs in a carbonate environment, acid leaching of the ore after beneficiation by radiometric sorting appears to be the most economic treatment method. Uranium would be recovered from the leach liquor by solvent extraction.

Jan 1, 1970

By Zichenko ZA, Volkov SV, Shaluhina LM, Emelyanov AF

A method of selective gold recovery from black sands has been elaborated which is based on the liquid phase low temperature chlorination of materials. A liquid agent mixture of sulphur monochloride S2C12 and sulphur dichloride SC12 is used. This method ensures the gold extraction either in the state of AuC13.SC14 complex salt solution or as metal. The kinetics and main physico-chemical parameters of the process (eg density, solid: liquid ratio, temperature) have been investigated. The materials investigated were quartz-insipid ride samples with low sulphur and high iron (up to 70 per cent Fe2O3) contents. Gold recovery is 96 - 99 per cent at temperatures not exceeding 100¦C. Most of the associated minerals remain in the cake. To obtain metallic gold, the resulting solution of the AuC13.SC14 salt after leaching is reduced by hydrazine with subsequent annealing. The elaborated method is remarkable for small energy expenditures owing to low process temperature and a high gold dissolution rate. The solution of sulphur chlorides and other reagents are used in closed cycle.

Jan 1, 1993

By Huntington J. F, Green A. A

The quality of a photointerpreter's anno- tation of an image may be displayed by writing maps in which the pixels are shaded according to the number of times a pixel was seen to be on a feature. The map of grey tones is more valuable, if, for each grey level, a measure of quality is determined, which is the probability that a feature exists in a pixel of a given grey tone. From repeated annotations (at least four) of the same image by a single observer, it is possible to formulate a perception model which describes stochastic processes in the observer. From that model, it is possible to assign estimates of data quality for each observer to each grey tone. For two observers on the same scene, the 'mean' method of quantifying the grey levels is to assign a value equal to the mean of the ii,:lities for each of the two observers separately. For this example, the grey tones arr thus quantified at probability levels of 0.70, 0.80, 0.90 and 0.95.

Jan 1, 1977

By Travis G. A

The Mons Cupri Cu-Pb-Zn deposit is situated within a complex Archaean volcanic sequence dominated by rhyolitic fragmental rocks. This sequence is underlain by a 150 m group of rhyolitic ash flow tuffs with minor conglomerate and sandstone, and is overlain by approximately 180 m of dacite and andesite flows. The Mons Cupri sequence itself consists of several elongate or circular rhyolite domes, flanked by rhyolitic fragmental rocks ranging up to 400 m in thickness away from the domes. The lithology of clasts (typically 1 cm to 1.5 m in size) provides strong evidence that the flanking fragmental unit was derived by explosive eruptions associated with the rhyolite domes. An approximately 15 m thick shale horizon, bleached and silicified at its base, overlies both the rhyolite domes and the frag- mental unit. A south-dipping chloritic pipe-like body, with near-surface dimensions of 1,000 by 250 m and narrowing at depth to produce a cone-shaped cross section, lies within the rhyolitic frag- mental rocks about 250 m north of the largest rhyolite'dome in the Mons Cupri vicinity. Its east west elongation parallels that of the rhyolite dome. Ghost fragments indicate that the material in this fissure is a severely chloritized derivative of the rhyolitic frag- mental unit. The chloritic pipe is overlain

Jan 1, 1973

By S Hancock, B H. Brady, H Muhlhaus

The aims of this paper are to present the governing science, verify the performance and demonstrate the potential engineering applications of a computational simulator for in situ leach mining of non-ferrous minerals. Assuming that the target mineral is deposited either uniformly within rock matrixes (matrix-hosted) or on fracture walls (fracture-hosted), in situ leaching can be treated as reactive flow and transport in geological media (porous media for matrix-hosted ore deposits and fractured media for fracture-hosted ore deposits). Under this treatment, the processes of in situ leach mining for non-ferrous minerals can be quantified through several sets of equations and these coupled equations have been numerically solved. The computational simulator has been applied to the simulations of one-, two-, and three-dimensional in situ leaching cases. The computational simulator, together with laboratory and in situ leaching tests, can be used to:optimise the leaching chemistry; optimise the in situ scale mining operations; scale up the in situ scale test operation to a full-scale mining operation; and evaluate the post-mining environmental restoration plan.

Jan 1, 2000

By T Ueda an, T Sasaoka, A Yabuki, H Shimada

The basic concept of auger mining is to extract coal beyond the economic limits of surface mining technology by drilling holes of an appropriate diameter size into the exposed seam of the highwall as deep as is technically, economically and operationally feasible. This method of mining is used at the final highwall of typical surface mining operations. This paper describes the auger mining systems and discusses the methods to increase the coal recovery in auger mining from a field trial in Australia.

Jan 1, 2003

With the recent increase in the number of underground secondary sulphide dust explosions in Canada, Australia and elsewhere in massive sulphide ore mining industry, establishment of quantitative and realistic methods of analyzing the mine dust is of paramount importance. This paper examines and analyses dust samples from a mine area that has very recently experienced sulphide dust explosions. Also, it attempts to evaluate the physical and chemical characteristics of the dust. Explosibility of the dust samples was determined by quantifying the oxidation reaction of the sulphide dust particles in a 1.2 L, Lindenberg open horizontal tube furnace. Percentage sulphur', % S oxidised during ignition of a dust sample of known assay, % S and dust concentration (g/m3) were determined by the quantitative chemical analysis of SO2 gas produced during the oxidation reaction. A dust sample with high sulphur content, 22.3% S (assayed) and mean particle size 9.82 pm, was found to exhibit a low degree of oxidation compared with a sample of low sulphur content, 15.44% S and mean particle size 21.7 tcm, at the same Temperature.

Jan 1, 1988

By Tilley H. L

The 90 m deep Yallourn Open Cut brown coal mine in the Latrobe Valley, Victoria is developing into an area close to a reverse fault. A major batter stabilisation program, matched to ongoing operations, has com- menced to maximise coal recovery where the fault is within 100 m of the Open Cut. The various mine development options considered during the planning stages are outlined. These options were formulated in a back- ground of geotechnical constraints and the need to maintain continuous and reliable coal supply at a time when a new coal field would be developed. The geotechnical investigations and analysis procedures under- taken to determine stability of the batters during their excavation are presented. The geotechnical constraints on the mine design and opera- tions required to ensure safe working conditions are detailed. Current operational and batter stability conditions are considered satisfactory.

Jan 1, 1989

By C D. Pocock

Maintaining catch berm capacity in open pit mines through controlled blasting prevents production delays associated with managing risks associated with wall instability. A method has been developed to quantify the drill hole deviation of presplit holes at Century Mine. Different controlled blasting techniques are used to control wall quality and decrease the risk. The performance of presplit drilling, in terms of deviation from design has been investigated. A methodology has been developed to quantify the drill hole deviation in three dimensions. This methodology was converted into a series of algorithms capable of handling an extensive range of data. The methodology was applied to a case study at Century Mine to evaluate the performance of drilling on two presplit blasts. Overall, 67 drill holes were investigated in the case study. The drill hole deviation was quantified in terms of three parameters: skewness, linearity and drill hole deviation parameter. Methods for quantifying each parameter were developed. Skewness, a measure of the amount of curvature and linearity, was defined as the measure of the consistency of the gradient of the drill hole. Drill hole deviation parameter is defined as the average measured deviation of the surveyed drill hole position from the design position. Acceptable drill hole deviation tolerances for each of the defined parameters were developed by analysing drill hole deviation parameters with respect to implications to catch berm capacity. Verification of the methodology was confirmed through the comparing of the completed drill hole deviation results using both visual inspection and cross analysis for each of the three methods. Holes with an R squared value of less than 0.97 for skewed deviation were classified as having skewed deviation, while holes with a difference in gradient of 3.5 were classified as having linear deviation. The results show that 76 per cent of the drill holes had deviation. Of this, 19 per cent were skewed, 57 per cent had linearity and 12 per cent contained both skewed and linearity deviation.

Nov 8, 2011

(Quality can be defined as that which is acceptable to the market and can be supplied with a predictable degree of uniformity and dependability at a low cost. The market for any mining product is downstream processing, whether it be an alumina refinery, a gold processing plant or a flotation plant. The cost of that downstream process is partly dependent on uniformity of its feed material both in physical and chemical terms. With more operations adopting procedures which require quality assurances in line with International Standards, from their raw material suppliers, miners must look to their grade control processes to provide uniformity and predicability.This is not a new concept. The mining industry has constantly been working with ahighly variable raw material supplied by Mother Nature, from whom we cannot demand quality assurances. Through the processes of exploration, resource and reserve definition, long and short term mine planning and day to day grade control an end product has been supplied which consistently meets our customers requirements, in an economic fashion. When any of these processes fail, then penalties are incurred which can range from increased mineral processing costs to loss of markets and in the ultimate case, failure of a mining operation.

Jan 1, 1992

Both the miner and the Regional Water Board have legal obligations to ensure that all discharges of water from a mine site are free of harmful wastes, and that proper authorisation is obtained to abstract, divert, or discharge water. This requires long-term planning of water management for the whole of the mine site. Guidance is given in the paper on the issues which must be covered in water right applications.

Jan 1, 1986