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By Hare G

The emphasis in Australia, until recently, has been on large scale mining methods with drive sizes in excess of 4.5 metres square and the acceptance of high dilution ratios, even when working in relatively narrow orebodies. The rising cost of extracting and treating the recovered ore and the lower prices available in the world markets has seen a growing trend to narrow mining methods.Traditionally narrow orebody mining has been carried out using hand held drilling methods and rail mounted boggers. Recent advances in mobile equipment now offers the operator of narrow method mines the opportunity to completely mechanise the mine from the drilling, through bogging to haulage, while maintaining minimal dilution in drive sizes to two metres wide or less.

Jan 1, 1993

By A L. Hemphill

Conventional models, based on pulp kinetics, were regressed directly to steady-state data from a nine-stage platinum flotation plant, to provide a basis for optimisation studies. æFroth factorsÆ were used to characterise the froth in each stage and a ædepressant factorÆ was used for gangue minerals, to account for addition of depressant to the cleaners. These factors improved the fit, but differences between plant and model of up to 50 per cent were still present in a number of streams in the cleaning circuit. A new model was developed, which incorporated two rate constants for the recycle of particles from the froth. These parameters provided a workable approximation of the distribution of probabilities of particles being recycled in plant cells, depending upon where particles appeared at the surface. The feed streams were characterised by two pulp kinetic parameters and the above recycle parameters, for all minerals and all stages. The fit to the data was improved very significantly, when compared to the use of conventional pulp-based models with several pulp rate parameters. This relatively simple model did not require additional data from batch tests and it was suitable for plant simulations.Simulations of various strategies for plant operation were tested, allowing the simulator to vary the concentrate flows from all or some of the stages, in order to maximise platinum recovery. The simulator was imbedded in an automatic optimiser and constraints on chrome content or mass flow of final concentrate were used. The simulations of the existing circuits demonstrated that an improvement in recovery, (for a constant concentrate mass) of about 0.5 per cent was possible, by optimising all concentrate flows. A significant reduction in the flow of scavenger concentrate and an increase in the recycle flows in the cleaners were noted. The results of simulations of various changes to the circuit configuration are also discussed. An increase in recovery of 1.4 per cent (for constant concentrate mass) has been achieved in recent simulations, through circuit changes. However, it was necessary to keep the mass pull of concentrate, per unit area, within typical limits.

Jan 1, 2005

By Fourney W. L, Weaver T. A

A large number of single borehole crater-forming shots were analyzed in terms of correlating the crater parameters with explosive weight. The objective was to determine a scaling rule for which the explosive weights varies over many orders of magnitude and covers many different geologic materials. Included in the data base are type and weight of the explosive, geologic medium and degree of saturation, charge depth-of-burial, crater dimensions, and crater rim (lip) measurements. An alternative scheme to the usual plot of scaled depth-of-burial versus scaled volume removed was desired. The proposed technique uses a plot of the logarithm of the explosive weight, Y, versus the logarithm of the slant length, R', calculated from the known depth-of-burial of the explosive and the measured crater radius. A plot of Log Y in kilotons versus Log R' in meters from all the crater tests in this data base resulted in a straight line with a of slope 0.31 and showed good correlation between the variables Log Y and Log R'. Since the crater data includes many geologic materials, different explosives (including nuclear tests), and large differences in charge weight, the straight line fit suggests that the type and strength of the geologic materials play a subordinate role to explosive energy in crater formation.\

Jan 1, 1990

By Delahey G, Dunne R

At New Celebration Gold Mine the advent of treating primary sulphide ores resulted in decreased gold recoveries of four to ten per cent. The grind coarsened due to an increase in ore hardness and gold in the residue was found to be associated with pyrite. Finer grinding improved recoveries but was impractical as production rates would have decreased to unacceptable levels. An alternative approach that has proved to be viable combines both gravity separation, to recover the coarse gold bearing sulphides, and then regrinding the concentrate in a tower mill prior to releaching in the cyanide leach circuit. The tailings treatment plant was commissioned in December 1990 and details of the first 18 months' operations are Provided.

Jan 1, 1993

By H Zahiri

Utilisation of the InSAR technique for monitoring of subsidence in mining areas, employing open pit and underground mining methods, has large potential due to sensitivity and safety issues associated with usage of the classical surveying techniques. InSAR can also be very competitive concerning the cost of provided results. However, there are a few issues that may significantly limit InSAR applicability for subsidence monitoring in mining areas. The altered terrain topography, involving steep slopes and deep pits, may lead to the layover of radar signal for specific satellite and pit geometry. Also, the highly dynamic character of subsidence induced by mining, especially using mass mining methods, may lead to issues such as apparent heave and ambiguous results. In this paper the authors analyse the above-mentioned issues and demonstrate how InSAR technology was applied, as a supporting system, to monitor large scale and highly dynamic subsidence for a real case study in Western Australia.

Jan 1, 2008

Warrego copper ore characteristically contains 0.05-0.2% bismuth, which on upgrading produces a concentrate of approximately 1.3% bismuth. Early smelting of this concentrate (1973) was accomplished in a flash furnace (eliminating 92% of the bismuth in the con- centrate) and Top Blown Rotary Convertors (eliminating 85% of the bismuth in the matte). However 1.3% bismuth was considered the max- imLun bismuth level that could be treated. Geological evidence indicated that the bismuth level in~the copper ore would increase and that the bismuth level in the Warrego ore would be 0.3-0.7% bismuth. Treating these ores would produce a copper concentrate with bismuth levels far in excess of those accept- able to the smelter and still produce low bismuth content blister (limit 0.03% bismuth).

Jan 1, 1984

Author Biography William Poole was born in Sydney, NSW in August 1868. He was educated at Sydney Grammar School. William spent two and a half years as a cadet in the Railway Construction Branch of the NSW Department of Public Works doing office and field work. He then undertook a civil engineering course at Sydney University. Following this he was employed for six years as assistant and resident engineer at the Water Conservation Branch, NSW and was in responsible charge of various field operations, surveys, river gaugings and works of constructions including a large timber weir and also a concrete weir on the Macquarie River, NSW. He then undertook a mining and metallurgical course at Sydney University and became a graduate (BE) with honours in civil engineering, mining engineering and metallurgy. He then spent four and a half years with BHP Co, one and a half years at Broken Hill in the mining and concentration works and three years at the smelting works in Port Pirie, SA, in the latter case he was metallurgist and night superintendent and during that time had responsibility and control at night of the blast furnaces and, during the last 15 months, of the whole of these large works. The author left the Port Pirie, SA works to assume the position of Director of the State School of Mines at Charters Towers and had responsible control of the establishment and also lectured in mining, surveying and structure. William Poole was honoured to hold membership in the following organisations: Assoc M Inst C E, L S of NSW, M Aus Inst M M, Assoc Inst M & M (London).

Jan 1, 2006

By P Gardner

The Kanowna Belle gold mine in Western Australia has reached a depth of 1000 m on the way to 1300 m. A feasibility study undertaken in 2006 identified the mining of E-Block as both technically and commercially feasible. One of the major risks identified in the study was the ventilation capacity, as the existing primary ventilation system would not support the production schedule for mining and a major system upgrade would be required. Failure to provide the necessary airflow would put the project at risk. A review of the airflow requirements and infrastructure options detailed in the feasibility study was undertaken to ensure that the assumptions and design were optimised. This review recommended a minimum airflow of 206 m3/s, but by acknowledging the necessary constraints imposed by mining at depth and re-engineering the mine layout to optimise the primary ventilation circuit, this requirement was reduced to 160 m3/s. This represents a 22 per cent reduction in airflow and a reduction in capital cost that will see mining at depth become economically feasible at Kanowna Belle. This paper discusses the processes that led to a redesign of the mining layout and highlights the importance of including the ventilation network in the optimisation process.

Jan 1, 2008

By Azizan A

The paper presents the results of chlorination of beneficiated rutile and struverite (a tantalum-niobium bearing, rutile type mineral in alluvial tin deposits) by carbon tetrachloride in the range 623¦K to 773¦K. The chlorination of rutile could reach near completion at 773¦K in about an hour. The process was found to follow a modified pore-blocking law, with the basic reaction constant K varying inversely with the particle size, so that the basic process may be a reaction under interface control (chemical control) as modified by the pore blocking effect. The pore blocking factors for the various experimental conditions were evaluated and the predicted curves showed good agreement with the experimental data. In the case of Struverite, a similar kinetic behaviour was observed but the chlorination process was relatively slower than that in rutile. It was also observed that in struverite, niobium got chlorinated faster, followed by tantalum and titanium, indicating a good possibility of separation by chlorination using carbon tetrachloride.

Jan 1, 1996

By Bailey S, Avraamides J

Laboratory-scale experiments were conducted in a mechanically stirred aeration vessel to determine the rusting abilities of aerated water in the presence of various catalysts compared to the industrially used ammonium chloride. The catalysts used were tartaric acid and citric acid, firstly by themselves, and then in later experiments mixed with ammonium chloride. From these experiments it was found that leaching of iron from reduced ilmenite in the presence of either tartaric acid or its sodium salt was much slower than in the presence of ammonium chloride. Citric acid or its sodium salt also give a slower rate of leaching than the ammonium chloride system, but citric acid and its sodium salt mixed in equal proportions produced a rate of leaching approximately twice as fast as the standard ammonium chloride system. When tartaric acid was mixed with ammonium chloride the leaching rate was found to be equivalent to the leaching rate of the same amount of ammonium chloride alone, thus indicating that tartaric acid made no contribution to the leaching in this svstem_

Jan 1, 1996

In the 100 years since the formation of The Australasian Institute of Mining and Metallurgy there have been significant changes in all aspects of the mining industry.A number of the changes have been driven by technical evolution or necessity, others have been driven by developments in the application of the products produced from the mining of minerals. Other changes have been driven by shifts in community values.

Jan 1, 1993

By Lo YC, Pate WT

In recent years there has been a great deal of interest in ways to optimise the performance of mineral processing plants. This paper describes a methodology developed for optimisation of autogenous and semiautogenous mills. The methodology involves plant sampling, batch testing, modelling, computer simulation, control strategy evaluation and implementation. The models have been tested in plant operations to verify the research findings and predictions. Two example applications of this approach to plant optimisation are presented. Benefits reported in these two examples are in the range of four - eight per cent increase in size reduction efficiency.

Jan 1, 1993

By Prakash AJ, Singh PK, Dhar BB

Safety and stability of underground mine openings, coal pillars and water dams in the near vicinity of operating surface mines are endangered from excessive blast induced vibrations. This paper deals with the investigations carried out in four coal mines. Simultaneous measurements at equal distances from the source indicated higher level of vibration on the surface than those produced in the underground workings. The values of peak particle velocity (PPV) measured on coal pillars, body of water dam were lower than those measured on roof and junctions of the roadways. The parting of coal (3-4.5 m) between two sections vibrated with higher PPV compared to the roof of top section and floor of bottom section. In one case the influence of total charge fired in a round was significant in generation of vibration compared to the maximum charge/delay where monitoring was done vertically below the source.

Jan 1, 1995

HISTORY OF PROCESS DEVELOPMENT The manufacture of coke from coal commenced at the beginning of the 17th century by heaping coal into circular mounds in the open, covering with earth, clay or ash and slowly burning the coal in the mounds to produce a heap of coke. Naturally the process was slow in production rate with low coke recovery due to the loss of that part of the coal which was completely burnt to ash. As a natural consequence of this initial carbonisation method, a coke oven design was developed so that the coking chamber, which was constructed of brick, was circular in plan, cylindrical in section with a dome shaped roof - similar in shape to the original coal mounds. As the shape of the chamber was similar to the old dome shaped beehive, the ovens were called beehive ovens. The ovens were built in long single or double rows with retaining walls between each oven and as such were known as a battery of beehive coke ovens. In these early beehive ovens, coal was charged from hand wheeled tubs into a hole into the top of each oven, hand- levelled with a rake and the oven side opening sealed with bricks and clay except for some small holes in the opening to admit limited quantity of air for combustion. Coking commenced by the heat retained in the brickwork from the previous charge. The gases evolved from the coal, together with the admitted air, were partially burnt inside the oven to supply the heat for carbonisation and the surplus gases were burnt to waste at the top charging hole. The heat from this gaseous combustion inside the oven together with the reflected heat from the brickwork of the arched roof of the oven allowed carbonisation of the coal charge to proceed from its surface slowly downwards to the floor of the oven. At the end of the coking cycle (usually 4 or 5 days) when there was no more evolution of gases, as indicated by diminishing gas flame and a drop in oven temperature, the brickwork of the oven opening was removed. The coke was then quenched by inserting a pipe with spray holes to spray water over the coke charge in the oven after which the coke was raked from the oven. The work of charging and levelling the coal, bricking and removal of oven opening, quenching and raking out the coke all required manual labour. As these ovens only carbonised a 5 to 7 ton charge of coal per cycle of 4 to 5 days, to produce 3 to 41/2 ton of marketable coke, the labour required per ton of coke was high.

Jan 1, 1975

By Campbell Brown G A

IN 1904-5 diamond drilling operations in the Mount Morgan mine proved the presence of some millions of tons of rich copper-gold ore, averaging about 45%to 50%silica and 22%Fe as sulphide. To smelt this ore in blast furnaces, without previous concentration, large quantities of basic flux were required. Unlimited quantities of cheap limestone were available at Marmor, on the GladstoneRockhampton railway, 50 miles from Mount Morgan, but the supply of iron flux was a more difficult problem. Smelting Operations started with one furnace in January, 1906, using as flux low-grade sulphide ore from the mine, containing very little silica. When the second furnace was blown in this was supplemented by ironstone from small deposits in the immediate district. It was recognized that these sources of supply would be insufficient to keep the full plant going after the third furnace had been blown in. Accordingly, the deposits at Kabra and Mount Etna were opened, neither of which proved capable of supplying the quantity required-viz., 1000 tons per week. Negotiations were meanwhile being conducted with the Marble Island Iron and Limestone Co. to obtain the right to mine at Iron Island, where a sufficiently large deposit was known to exist. Shipments from this source commenced in January, 1907, and ceased at the end of July of this year, when the metallic flux from...

Jan 1, 1911

By Nguyen G. H, Graham J

A laboratory-scale fluidised-bed roaster has been used to examine the effect of temperature and duration of roast on the morphology of the calcine produced from a pyritic gold-bearing concentrate. Two mechanisms of oxidation, which depend on the roasting temperature, have been identified. For samples roasted below 550°C, direct oxidation of the sulphide occurred to form hematite with a porous fine-grained concentrically-layered texture. The surface area of the calcine increased with an increase in roasting temperature in the range 400 to 500°C. For samples roasted above 550°C, the first reaction was the evolution of sulphur vapour to produce pyrrhotite, which underwent increasing grain growth as the roasting temperature was increased. This had the effect of decreasing the surface area of the calcine. The pyrrhotite was subsequently oxidised to produce hematite.Leaching studies using cyanide solution indicated that gold extraction around 95 per cent was achieved at temperatures of 500°C and above, with optimum extraction, taking into account rate and cyanide consumption in the range 550 - 700°C. Hence recovery did not decrease with decrease in surface area of the calcine. Observed gold in leached calcines was mainly enclosed in thin oxide walls of the hematite, with minor quantities enclosed in quartz and silicate minerals.

Jan 1, 1995

By Trutwin W, Tracz J

The outbreak of open fires underground cause disturbances to the conditions of the mine ventilation process. Thermal and chemical effects due to fire create unsteady, transient states in airflow and the spread of combustion products in the mine ventilation network. Changes in airflow have an influence on the fire, constituting a sort of feed-back between the fire and the ventilation process. This paper deals with a numerical simulation method, which enables the determination of transients caused by fires. The mathematical model, of the simulation method considered, has been based upon a: simple model of combustion of fuel in the airway, unsteady exchange of heat between the fire, airflow and surrounding rocks, airflow in mine ventilation networks and the theory of heatflow in the rocks. Initial and boundary conditions allow the determination of a set of parameters describing the airflow, the propagation of combustion products, the tempera- ture of the fire and flow etc. due to the fire. The simulation method, facilitates a graphical presentation of the spread of combustion products throughout the network. The simulation method has been prepared for an IBM PC/AT Computer.

Jan 1, 1988

The alluvial diamond deposit in the Kimberley Region of Western Australia which became to be known as the Bow River Diamond Mine was discovered in 1982 by joint venture partners Freeport of Australia Inc and Gem Exploration and Minerals Ltd. Mining commenced in 1987. This paper describes the exploration programs and results from discovery to commencement of mining and reviews the public reporting of this process in the light of 1992 Ore Reserves Code (JORC, 1992). This case study illustrates the difficulties involved in resource estimation of alluvial diamond deposits. The study also shows that current reporting standards were, to a large extent, adhered to a full ten years before the diamond reporting section to the Code was formulated.

Jan 1, 1996

By Hutchings AA

This paper describes MIM's experience in arranging and executing project financings for coal projects at Oaky Creek and Newlands and Collinsville and the port facility of Abbot Point.The financings had their origins in the late 1970's and culminated with the opening of these projects during 1983 and 1984.MIM's financing developments therefore span an important period in the development of resource projects in Australia.Resources projects have been growing in both scale and complexity of financing requirements.

Jan 1, 1985

By Askew JE

The large, massive low grade gold orebody at Mount Charlotte has demanded an innovative approach to mining engineering. In the last five years several advances occurred in the areas of rock mechanics, mass blasting, blasthole drilling and truck haulage.Successive crown pillar failures throughout the life of the mine had resulted in a continuous geotechnical program, mostly utilisingconsultants. The culmination of this workwas the development of a successful crown pillar support system. Several stopes have incorporated the technique and the most notable success has been the relatively large F3 crown pillar.In hole hammer drilling was introduced to the mine in 1979, but little work had been done to develop a blasting system to suit the 165mm diameter holes. A non-electric sequential blasting system was adapted for use at themine and up to 600,000 tonnes have been fired successfully in an individual blast. Problems with ground vibration were addressed although this work remains incomplete.

Jan 1, 1984