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By Khosia NK, Ravishankar SA, Sankar TAP

Alumina reduction in Rajhara iron ore slimes essentially involves separation of hematite from gibbsite in relatively coarser fraction and from kaolinite in the finer fraction. Selective dispersion flocculation studies on synthetic mixtures clearly indicated that while it was possible to achieve almost perfect separation of hematite from kaolinite using starch and PAMX (modified polyacrylamide containing hydroxamate functional groups) flocculants, no selectivity was observed in iron oxide-alumina system. A schematic flowsheet suggested on the basis of this work for the beneficiation of Indian ore slimes thus consists of classification in a hydrocyclone, followed by selective dispersion of coarse fraction (underflow) and selective flocculation of the overflow fraction. Starting with a feed assaying 46 per cent Fe and 11.6 per cent A1203, a concentrate assaying 60 per cent Fe and 4.7 per cent A1203 could be produced at an overall iron recovery of 60 per cent using conventional Reagents.

Jan 1, 1993

By Kokubu H, Moro-Oka A

Gaseous desulphurization from hot metal- lurgical slags containing sulphur mainly as sulphide and(or) sulphate was measured under `low at four different oxygen pressures. The total gas rate was from 1.7 to 2,8 litre/min. Crushed slag samples weighing 2 to 12 g were used for the experiments. The amount of SO2 gas evolved from samples kept at 15500 was rieasured over tires ranging from 20 to 300 min. The gaseous desulphurization rate of sul- phide sulphur becomes high as the oxygen part- ial pressure is increased, while that of sul- phate sulphur is high under low oxygen pressure. The effect of basicity on the desulphurization rate of sulphide sulphur in oxygen atmospheres is not obvious in the range of basicity from one to three. Changes in the ratio of sulphide to sulphate sulphur during the run in oxygen axe analysed. Desulphurization reactions and the rate controlling steps are discussed. De- sulphurization phenomena can be explained by inte.-¦ucia.l chemical reaction rate control.

Jan 1, 1980

By Wilson J. R

The main mining method employed at Mt Lyell is sub-level open stoping with a typical extraction level interval of 100 m and sub-level intervals of 30 m. Currently, pillar blasts provide the major portion of broken ore and firings are in the order of 200 000 to 300 000 tonnes while future firings of 500 000 to 300 000 tonnes are planned. Improvement of blasting techniques in these mass firings has been necessary for the following reasons. 1. Extremely wet ground conditions have adversely affected ANFO performance and placed extreme pressure on the charging and priming phases. 2. As pillar blasts increase in size the problems of wet ground conditions are accentuated resulting in very slow loading of wet upholes with cartridged gelignite and unreliable charging of the remainder with ANFO. 3 Fragmentation in pillar blasts has suffered because of the above, result- ing in increased secondary breaking costs and difficulty of extraction.

Jan 1, 1977

By Olofsson O, Durance M-V, Villeneuve J

Modelling of the Sala Agitated Mill (SAM) using a simplified solution to the population balance approach has been carried out. The rate of breakage is modified by the mill power, providing predictions of the resulting product size distribution for different grinding conditions and mill sizes. The power of the mill is estimated by a physically derived function, using an assumption that the mill power consumption mainly arises from the ball to ball friction within the mill. The grinding efficiency, depending on the media ball size and pulp density is approximated in the model. Experimental work on grinding a complex sulphide ore using a pilot scale SAM 7.5 has been carried out. Good agreement has been found between the data from the test campaign and the simulated values. The model has also been validated with experimental data from a pilot grinding of calcium carbonate and from an industrial mill. The model is included in the USIM PAC simulation software.

Jan 1, 1995

The Beaconsfield Gold Mine within the township of Beaconsfield in northern Tasmania has been established in recent years to exploit a known resource in the Tasmania Reef which was mined to a depth of 450 m below surface between 1877 and 1914 yielding 850 000 ounces of gold. Activity at Beaconsfield over the past 30 years has involved drilling to assess the resource and rehabilitation of the top 100 m of the Hart Shaft that had collapsed since the original closure. More recently the mine has been dewatered to the 375 m Level and work is well advanced establishing infrastructure aimed at recommencing production during 1998 - 99. This paper reviews geotechnical studies undertaken to-date and provides a brief overview of some practical methods of analysing ground conditions. It pursues simple observation and analysis as a tool for assessing mine design and ground support options without recourse to extensive monitoring and measurement and avoids the use of elaborate models before the basics are understood.

Jan 1, 1997

By Elford H. S, Hughes J. F

The object of this paper is to bring before Members of the lnstitute the very important and much discussed question of scientifically prospecting for ores, oil, coal, etc.The paper is divided into three sections:SECTION l - Description of Geophysical Prospecting methods.(1) Sound-Vibration.(2) Magnetic.(3) Gravitometric.(4) Electrical.(a) Equipotential.(b) Electro-magnetic.(c) Schlumberger.(d) General.SECTION II - General discussion of Cicophysical Prospecting.SECTION III - Appendices.(1) Systematic summary of methods of survey.(2) Table of electrical conductivities.(3) Bibliography.SECTION I - DESCRIPTION O:F GEOPHYSICAL PROSPECTING METHODS.(1) SOUND-VIBRATION.THE transmission of sound waves or vibrations underground will depend on at least three factors: (a) Type of Rock. - Hard, dense rocks wiII transmit sound waves much more quickly than softer, more porous rocks. In the former the speed may be eight times that of the latter.(2) Uniformity of Rock. - Any variation in the texture, hardness, etc, wiII have an effect on and direction of transmission of the waves.

Jan 1, 1927

By Frith RT, M D. Goodz

Drilling has been accepted as a major exploration tool. However its use in mine planning and development has been overlooked.This need for drillhole data was highlighted during the redevelopment boom of the central Victorian gold mines. During the 1980's numerous old mines were evaluated and successfully drilled to yield mineable parcels of quartz reefs close to existing workings.Invariably the initial decision was always to attempt reopening and rehabilitating existing access. The four case studies described have proceededsufficiently to show that major cost savings would have been made if underground drilling had been used as the main tool in mine planning.Savings not only in expenditure, but in the time required to make key decisions. In two cases the mines would have experienced far earlier expansions in production. In one case the mine would be currently producing rather than awaiting further finance. In the fourth example, the decision to case development would have been much earlier providing capital to other projects.

Jan 1, 1990

By Thomas MJ

The Cannington silver-lead-zinc deposit is located at latitude 21¦ 52' 9" South, longitude 140¦ 55' 10"" East, on the south-eastern margin of the Proterozoic Mount Isa Block (Figure 1), 135 kilometres south-east of Cloncurry. Access is via a graded shire road for 80 kilometres south from the small pastoral community of McKinlay.REGIONALThe deposit is hosted by the Proterozoic Soldiers Cap Group within the Mount Isa Block Eastern Succession. South of the Cloncurry Beardsmore et al (1988) have grouped part of the Eastern Succession stratigraphy intothe Fullarton River and Soldiers Cap groups (1760-1740 million years) which together make up the Maronan Super Group. This is an apparently conformable sequence 10 kilometres thick and consists of metamorphosed sediments and volcanics with intrusive amphibolite bodies.The Maronan Super Group has experienced a complex history of deformation with up to six events having been recognized. The stratigraphy trends broadly north-south and is predominantly characterized by tight, upright isoclinal fold structures.

Jan 1, 1993

By Lauha E. A

The Goldstrike Mine is in Eureka County, Nevada, U.S.A., approximately 23 miles northwest of the town of Carlin. The mine is at the north end of the Tuscarora Spur of the Tuscarora Mountains at elevations of 5,300 feet to 5,950 feet. It is part of the Bluestar/Goldstrike subdistrict of the Carlin Trend, a northwest alignment of sediment hosted, disseminated gold deposits. The Carlin Trend includes over 15 major gold deposits with production, reserves and geological resources of approximately 65 million ounces of gold (Figure 1). In the Bluestar/Goldstrike subdistrict, there are seven major deposits; Betze, Lower Post, Post, Genesis, Bluestar, Purple Vein and Deep Star, plus numerous smaller deposits (Figure 2). Production, reserves and resources for the subdistrict are approximately 40 million ounces. Betze and Purple Vein are located on property controlled by Bamck Goldstrike Mines and Bluestar and Genesis on Newmont Gold property. Post, Lower Post, and Deep Star occur both on Barrick and Newmont Property.

Jan 1, 1991

By Rowley M, Bradford P

Plutonic Resources Limited (formerly known as Pioneer Mineral Exploration Limited) delineated, developed and operates the Plutonic Gold Mine located 180 kms northeast of Meekatharra, Western Australia. Gold mineralisation was discovered in the Marymia greenstone belt by Great Central Mines NL (GCM) in 1988. There was no previous, recorded gold mining activity in the area. In May 1989, Plutonic Resources Limited (PRL) entered into an agreement to acquire a 51 per cent interest in the three Mining Leases containing the Plutonic deposit from GCM. Subsequently in December 1989, PRL acquired the remaining 49 per cent interest in the Mining Leases. PRL and GCM are also involved in a 51:49 regional joint venture on extensive Exploration Licences in the vicinity of the Mining Leases. Following due diligence validation, reserve delineation and testwork, a feasibility study was undertaken in conjunction with JR Engineering Services, which led to the financing and development of the project for operation in August 1990.

Jan 1, 1991

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 T Gibbons, M Korte, C Brent, K Runge

The Barrick Osborne concentrator has gone through a number of upgrades since commissioning and steadily increased production from the original 119 t/h design capacity to 265 t/h milling throughput in 2008. This was achieved incrementally over several years, by maximising operating efficiencies, in order to minimise capital expenditure, and take advantages of economies of scale; but more importantly it was reached with no measurable loss of recovery. Osborne currently processes a blend of under- ground with open pit material from the nearby Trekelano deposit. Following primary/secondary crushing, the material is processed through a rod/ball mill grinding circuit ahead of rougher/scavenger flotation that produces a copper concentrate.Osborne was seeking a further increase in plant throughput of the order of 15 per cent, but at more sustainable levels of secondary crusher plant utilisation. In early 2008, Osborne contracted Metso Minerals Process Technology Asia-Pacific and South America (MMPT) to review the entire process from run-of-mine feed to final copper concentrate. This review would be done concurrently, with surveys of both the crushing/ grinding and flotation circuits being performed at the same time. The outcome would be a steady-state model of the complete flow sheet allowing the crushing/grinding model to interact with the flotation circuit model such that overall process optimisation was possible. Two key outcomes for the operation are exploration of the operating limits of the currently installed plant, and projection of the logical next stage of potential capital plant expansions. This paper covers Part B or the flotation component of the project. A separate paper discussing Part A or crushing and grinding is also being presented at this conference. Surveys of the Osborne flotation circuit were performed along with a comprehensive review of the cell characteristics including measurements of the gas and froth phases of individual flotation cells. These data resulted in a size-by-size floatability component model of the complete circuit. This circuit model was then used to perform steady-state simulat- ions to evaluate options to improve both overall circuit and individual cell performance. The grinding circuit model discussed in a separate paper or Part A of this project was then used interactively with the flotation circuit model to recommend the best operating regime to maximise overall efficiency. For example, the balance between increased crushing and grinding costs versus higher copper recovery and final concentrate grade. Due to changes in market conditions in the fourth quarter of 2008, the focus was shifted to prioritising operational changes rather than capital expansion modifications, in order to improve circuit performance.

Jan 1, 2009

The West Fork project is located in Reynolds County, Missouri, in the mineralized zone referred to as the Viburnum Trend. The West Fork orebody has a calculated ore re- serve of 13 600 000 tonne (15 000 000 st) con- taining 5.5 per cent lead, 1.2 per cent zinc, and recoverable values of copper and silver. A total of 64 600 tonne (71 200 st) of lead concentrate and 11 300 tonne (12 500 st) of zinc concentrate will be produced annually. The mill was designed to operate at an average capacity of 3450 tonne/day (3800 st/day), based on a 95 per cent availability factor. The mill will operate 24 hours a day, 5 days a week, with maintenance performed, as needed, on weekends. The mine will operate on an identical schedule; but the underground, pri- mary crushing will operate on a six hour/shift, three shift/day schedule. The 1.07m x 1.22m (42in x 48in) jaw crusher will produce a minus 20cm (8 in) material at an average rate of 191 tonne/h (211 st/h). The jaw crusher has almost twice the throughput capacity required, but the decision to install that size of crusher was based on the need to handle over sized rock in the ore.The coarse crushed ore will discharge into a 180 tonne (200 st) skip pocket, which will allow up to a one hour surge capacity prior to hoisting. Hoisted ore will discharge into a 900 tonne (1000 st) storage bin located adjacent to the headframe. The skip pocket and coarse ore silo comprise the total ore storage capacity. The concept of having no intermediate fine ore storage between the secondary crushing and grinding sections was borrowed from the Brushy Creek Mill of St. Joe Minerals, which operates adjacent to the West Fork property. Integration of the secondary crushing and grinding circuits has worked ' successfully for years at the Brushy Creek Mill. The most obvious advantage to eliminating the intermediate fine ore storage is'capital expen- diture savings. The capital expenditure savings include the cost of silos, feeders, and the cost of additional conveyor length and eleva- tion. Savings in operating costs will also be gained, particularly in reduced maintenance costs. Other factors which influenced the decision to eliminate fine ore storage were the simplicity of the secondary crushing sec- tion and the limited area available for con struction of the facility

Jan 1, 1978

MR JOHN S Foxall wrote: Mr Clayton in his paper on the above subject (Proceedings, N.S. No. 34) made an assertion which should not be allowed to remain unchallenged. On the top of page 10 appears the following statement;- "By viewing the box at adepression angle of 45°, all the dimensions are disclosed and the box may be projected on isometric co-ordinates" and lower down on the same page:- "An isometric plan, being virtually a projection on a plane inclined at 45°, the nearer the dip approaches this figure so is the outline of the stoping more truly shown." Those statements embody an inaccuracy which would be misleading were an isometric plan to be constructed to suit a special case, ie the case of a mine whose lode dipped at 45°, or approximately 45° and it would be found that in such a plan the outline of the stoping would not be truly shown. The reason for this is that the depression angle at which an object is viewed in order to...

Jan 1, 1920

The report by Baker (1958, p. 27) of krennerite in contact with gold in a specimen from the 2500 ft level, Phantom South lode, Great Boulder mine, Kalgoorlie, led the writer to look into this further, since Cabri (1965) reported that the assemblage krennerite-gold was unstable between 238° and 356° C, and possibly down to room temperature.Dr. G. Baker very generously provided the writer with two chips from the same rock specimen (No. 14 as labelled from the mine) in which the krennerite gold assemblage was reported to occur. These were examined by polished section and x-ray diffraction techniques and returned to the classified collection of the Mineragraphic Investigations Section of CSIRO.One specimen, showing visible gold, was examined optically with a binocular microscope and telluride material was twice removed from contact with the gold for x-ray diffraction. This material proved to be petzite and so amends a statement by Baker (1958, p. 39) that no petzite was found in the Phantom lodes.The second chip was polished, and microscopic...

Jan 1, 1967

Case Study: Dragline Swing to Dump Automation By Peter Corke, CSIRO Manufacturing Technology/CRC for Mining Technology and Equipment (CMTE) Peter Corke presented a case study of a project to automate the dragline swing to dump operation. The project is funded by ACARP, BHP Coal, Pacific Coal and the CMTE and is being carried out on a dragline at Pacific Coal's Meandu mine near Brisbane. Corke began by highlighting that the minerals industry makes extensive use of large, mechanised machines. However, unlike other industries, mining has not adopted automation and most machines are controlled by human operators on board the machine itself. Choosing an automation target The dragline automation was chosen because: ò draglines are one of the biggest capital assets in a mine; ò performance between operators vary significantly, so improved capital utilisation is possible; ò the dragline is often the bottleneck in production; ò a large part of the operation cycle is spent swinging from dig to dump; and ò it is technically feasible. There has been a history of drag line automation projects, none with great success.

Jan 1, 1997

Dynamic hydrothermal modeling is a tool that can aid in studying problems associated with hydrothermal ore deposits and can help increase our understanding of the spatial distribution of ore within hydrothermal systems. Models are constructed by a sequence of steps involving development of a conceptual model, followed by its mathematical description in terms of governing equation with accompanying initial and boundary conditions. This equation is then solved either analytically or numerically. The four model types ò that fully describe a hydrothermal 'system are those for fluid flow, heat and ore solute transport, and deformation. At Renison, models have been used to help understand two major exploration problems. These are the controls on ore localisation in dolomite horizons and feeder faults respectively. The types of models used ò have simulated the distribution of all mineral zones, pyrrhotite, and temperature in the dolomites. This has allowed evaluation of the effects on ore localisation by physical and chemical heterogeneities within the dolomite host rocks, and allowed estimation of potential controls in different physical and chemical environments.

Jan 1, 1987

The Newmont Waihi Operations mill has undergone significant growth over its life. Initially designed as a single SAG mill operation, the circuit has been progressively upgraded since 1988 from a nameplate capacity of 600 000 tonnes per annum to the current rate of 1.3 million tonnes per annum. With future mining potential there is renewed interest to increase the processing rate even further. Over the next twelve months five sustaining capital projects will be undertaken and two major eco-efficiency projects will be completed. With the exception of the large grinding mills, most equipment on a gold treatment plant is replaced after 10-15 years. Waihi Gold mine has been operating for 14 years and has just recently received approval to replace equipment in five main areas; carbon elution, carbon regeneration, scats crushing, intertank screens and variable speed drives. Each project has been justified on improved process efficiencies and operating/maintenance cost savings. The two eco-efficiency projects involve a water balance model and a cyanide optimisation project, both of which are Corporate key performance indicators (KPI).

Jan 1, 2002

The mining industry represents a small part of the international economy. Comprising few large companies, the industry exhibits volatile returns and over the past 25 years has failed to produce a long-term return that meets its cost of capital. It is not surprising, therefore, that the minerals extraction and processing industry is under considerable pressure from the investment community to lift its game. Management and boards are responding to this pressure by implementing more exacting cost control measures, capital management programs and through the process of consolidation. However, the cost of capital to the industry appears to have increased in recent years and investor attention seems to be preoccupied with the ænewÆ economy. Both of these influences will conspire to raise the bar for acceptable industry economic performance. This paper provides a useful snapshot of the minerals and extraction industry as it enters the 21st century. It remains cautious about future performance other than to conclude that single digit returns are likely to prevail, albeit at higher levels than over the past 25 years.

Jan 1, 2000

By Jefferson CB

Adaptive optimisation is the adjustment of operating conditions to maintain optimum process performance as the feed quality varies. This strategy has the potential to be economic- ally better than the maintenance of a constant optimum performance with a blended feed. Application of the strategy is hindered by practical difficulties and economic risks. A computer simulation of leaching processes has been used to identify ore-process systems which have the greatest potential benefits from adaptive optimisation. They are systems with low value feed materials and a broad distribution of reactivities, in pro- cesses which are sensitive to the operating variable that is being manipulated (e.g. residence time). Improvements of up to 13% in throughput and 5% in the economic operating margin can be achieved. The results are sen- sitive to the precise model which is used for the process kinetics. It seems that the kinetics of the last few per cent to react are particularly relevant. A realistic situation has been identified where a synergistic inter- action occurs in blended feeds which makes adaptive optimisation disadvantageous.

Jan 1, 1984