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By Wright D. J

Successful heap leaching depends on three major factors: 1. Chemical conditions appropriate for the dissolution of the valuable metal; 2. Contact between the leach solution and the valuable metal particles; and 3. Movement of the leach solution through, and out of, the heap. Given that appropriate chemical conditions can be provided, and the individual particles have sufficient permeability to allow solute/solvent contact, the problem of getting the solution to and from the particles remains. Non-uniform percolation of solution throughout the heap, caused by the presence of fines, is a common problem which occurs, to some extent, with all heap leachable ore. The conventional solution to this problem is agglomeration, which involves the attachment of fine particles to larger particles, effectively modifying the size distribution by preventing the fine particles from acting on their own. While agglomeration must be considered as an extremely effective technique with virtually no limit to its application from a technical point of view, its application is limited by economic considerations. A new technique has been developed which is not only a low cost alternative to agglomeration, but also has the potential for enabling the application of low cost heap and vat leach techniques to a significantly wider range of ores than current practice would suggest. The new technique, known as Electrokinetic Leaching, utilises electro-osmotic flow to enhance percolation through the heap. This paper describes and presents results of the laboratory testing, pilot heap operation and commercial application of Electrokinetic Leaching of a low grade gold ore and a tailings deposit. Comparative cost analyses between conventional and the alternative processing techniques are presented.

Jan 1, 1994

By A. W. CAMERON

Over a 4 hr period of steady plant operation samples were taken at regular intervals from points within the grinding and flotation circuits. Size fractions of each composite sample were analysed for lead and zinc values and a comparison made of the behaviour of the lead, zinc, and gangue minerals, size by size throughout the whole circuit.Analysis of the breakage behaviour using dynamic mathematical models gave the rates of breakage of each size fraction of each main component. Simulated grinding experiments showed that the main grinding circuit had some reserve capacity.In the lead flotation circuit all lead values between 150 mesh (Tyler) and 5 microns had excellent flotation characteristics, and a major proportion had a high rate of flotation which was practically the same for all sizes. Increase in overall lead recovery would require improvement of the flotation behaviour of ultra-fine lead minerals. These were slow floating and were responsible for approximately half the lead losses in the final zinc concentrate and one-third of the lead losses in the final tails.All sizes of zinc values between 100 mesh and 7 microns floated well and the performance of the zinc flotation circuit was mainly limited by the relatively poor flotation of the very fine zinc values. Approximately 60 per cent of the zinc lost to the final tails lay within the minus 7 micron fraction.The overall flotation conditions were conducive to very good selective flotation of the major proportion of the values and limitations in metallurgical performance were mainly due to the properties of the ultra-fines.

Jan 1, 1971

It is recognised by the current JORC (1999) code that resource classification involves the interaction of numerous qualitative and quantitative criteria such as data quality, geological continuity and grade continuity. No prescribed criteria and rules will work for all situations or even between different ore types within the same deposit. Also these criteria and rules are sometimes applied without a clear understanding of their appropriateness, accuracy or correct implementation. For these reasons case studies are useful to evaluate and compare criteria commonly used to assist in resource classification. In this paper blasthole data from a selected area of Joffre Member hosted ore of the Brockman Iron Formation at the Mt Whaleback orebody are used as the basis of a case study for the above-mentioned purposes. Iron grade, in per cent, is interpolated from a blasthole dataset into a block model using ordinary kriging. Samples are then removed from this blasthole dataset to produce a random sample grid, a semi-random sample grid and a regular sample grid. Using nearest neighbour, inverse distance squared, ordinary kriging and sequential gaussian simulation, block iron grades are estimated using these three blasthole data subsets as inputs. This provides four groups of estimates with the ordinary kriging estimate based on the complete blasthole dataset being considered to represent the true estimate of iron grades. Also generated during this process are estimates of the error for each block. These error estimates range from the simple, such as drill spacing, through to more advanced methodologies such as kriging efficiency and simulation. These estimates of grade and grade estimation error are then compared to the true estimate using graphs and statistics. In order of increasing accuracy the block grade estimation and simulation methods were nearest neighbour, inverse distance weighting, ordinary kriging and finally conditional simulation. The simulation in some instances doubled the accuracy of individual block grade estimates when compared to nearest neighbour and inverse distance weighting estimates. It was found that many methodologies for estimating the error of individual blocks performed equally well with only methods such as average drill hole spacing and classification by search ellipse pass number performing poorly. An approach on how to convert the abovementioned error estimates of individual blocks into a meaningful JORC classification is also discussed. Although advanced non-linear resource estimates are applicable, most iron ore mines are still using relatively straightforward methods. The use of blasthole data and some simple linear interpolation methods and simple linear based error estimates makes this study repeatable for most iron ore sites and their resource geologists. This style of investigation is recommended as a useful approach for the Competent Person to apply to their deposit and thus better select, implement and understand the criteria used for resource classification and provide more consistency and confidence in the resource classification process.

Jan 1, 2002

During the past decade geopolymer technology has emerged as a possible solution for the effective stabilisation and solidification of waste materials. Amorphous to semi-crystalline geopolymers form by the co-polymerisation of individual alumino and silicate species, which originate from the dissolution of silicon and aluminium containing source materials at a high pH in the presence of soluble alkali metal silicates. It has been shown previously that geopolymerisation can transform a wide range of waste alumino-silicate materials such as fly ash and blast furnace slag into building and mining products with excellent chemical and physical properties, including high compressive strength as well as fire and acid resistance. These chemical and physical properties suggest that geopolymeric matrices derived from waste materials are well suited for environmental and other applications such as the encapsulation of mine tails, the immobilisation of heavy metals and also the fabrication of structural products. The current challenge is to demonstrate that a wide variety of natural Al-Si minerals could serve as potential source materials for the synthesis of geopolymers. Moreover, it should be shown on a large commercial scale that it is possible to encapsulate waste materials into the geopolymeric structure without compromising the chemical and physical properties of the final product for the desired application.

Jan 1, 2000

The seventies was a decade of intensive hydrometallurgical activity, firstly for copper recovery, then uranium, and finally gold. In comminution, semi-autogenous grinding was applied to uranium and gold ores. In leaching there were extensive investigations into pressure and in-situ systems, and into ammonia and chloride leaching processes. Solvent extraction became a major unit operation for copper recovery. New reagents and equipment were introduced. There was a resurgence of ion exchange for uranium as low grade sources were tapped and continuous ion exchange systems developed. Carbon-in-pulp technology came to the fore for gold recovery with the development of desorption and electrowinning systems. Electrowinning was widely applied for copper in conjunction with solvent extraction. There was a continued move to greater mechanisation and higher current density. There were important developments in solid-liquid separation such as high capacity thickeners, belt filters and sand filters. Many of these trends will continue in the eighties. Interesting new developments can be expected.

Jan 1, 1981

THERE is available a considerable amount of literature concerning the practice, and, to a smaller extent, the theory of ventilation as it applies to coal mines. Publications considering the subject of ventilation of metal mines from a practical viewpoint are much less numerous, whilst the treatment of the theoretical aspect of the question (except in so far as is common to coal mining) appears to have been almost neglected. This may be due to one of two causes (a) That in many metal mines (especially those which have not attained a great depth), on account of the greater number of connections with the surface, as compared with coal mines, natural ventilation has sufficed.(b) That in metal mines where the foregoing is not the case the solution of the ventilation problem from theoretical considerations has been regarded as involving so many complications that any attempt at the determination of a set of concrete ventilation figures applying to and throughout the mine would be futile. Whilst it is freely admitted that the difficulties inherent in the application of theory to the ventilation of metal mines are numerous, it must also be admitted that few attempts have been made to overcome them. The ultimate aim, of course, would be the collection of sufficient data to enable the probable behaviour of the air-currents throughout the whole of the workings of any mine to be accurately predicted under any particular set of circumstances. Failing the attainment of this goal of perfection, it would at least appear that much valuable information might be obtained. In ventilation, as in...

Jan 1, 1921

The medical evaluation of the health effects of dust exposure in the mining industry should be based on the premise that the results of the medical examinations will be considered in relation to the previous history of exposure to dust.(l) To this end the Mines Regulation Act and the Regulations under that Act in West- ern Australia provide for: the prime responsibility for dust sampling to rest with the mining companies; the appointment of a Ventilation officer by each mine and the sampling and record- ing by him of dust and toxic gases in work places at three monthly intervals, or more frequently as required; the appointment of a Mines Medical Officer who shall make provision for the medical examination of persons and the prevention of occupational diseases; the appointment of a Ventilation Board, consisting of the State Mining Engineer, the Senior Inspector of Mines for the State, an Inspector of Mines having specialised training in ventilation tech- nology, a medical practitioner having specialised experience in occupational health problems relevant to the mining industry, and a Scientific Officer having duties under the provisions of the Clean Air Act, 1964;

Jan 1, 1979

The CSIRO acid alumina employs the essential steps of digestion, hydrolysis and calcination to produce pure alumina from ores too high in silica or too low in grade for the Bayer process. The two latter stages have been described in some detail previously but the digestion stage requires separate dis-cussion since ores vary widely in composition and reactivity. Kaolinitic clays require. roasting to 8000C to render the alumina soluble in the 11 per cent sulphuric acid used in the process while some ferruginous laterites require roasting to reduce the reactivity of iron minerals. The iron content of pregnant solutions is kept at a low figure by digesting at temperatures where the solubilities of ferric oxide or basic ferric sulphates are small. A.two-stage, counter-current digestion is necessary to extract a high percentage of alumina from the ores and to pro-duce a basic pregnant liquor, with an S03:Al203 ratio of 1. 9 suitable for high temperature hydrolysis. The second stage of digestion, called the modification stage, treats slightly acid liquors with exces's of ore in the temperature range 1200- 1400C. The first stage operates at 1700-190 C to extract further alumina from modification residues, using recycled acid obtained from later steps of the process. Both temperature ranges are critical for economic operation and for proper con-trol of the iron content of solutions. In batch experiments, thirty minutes is regarded as a suitable retention time for digestion and modification; ores which require much longer than this are less suitable for the process. Impurities. such assilica, titanium and vanadium oxides are discarded in the residue along with iron oxide and do not contaminate the final product. Bivalent metallic sulphates such as MgSO4 can build up in the circuit to a very considerable degree without affecting. product purity.Comparison with other processes shows that the C. S. I. R. O. process is superior for control of contamination by iron, which is the major problem in acid processes. The pure intermediate product (3 A1203. 4 S03. 9 H2O) is also more convenient for cal-cination to alumina than the hydrated aluminium sulphate usually recovered in other processes. Detailed cost comparisons must await pilot-plant evaluation of the C. S. I. R. O. process.

Jan 1, 1963

During 1995 a catastrophic accident took place at the No 2 Vaal Reefs shaft in South Africa, when a locomotive fell onto a cage, causing the death of 104 people. The subsequent Commission of Enquiry found that the safety detaching hook had opened when hit by parts of the falling locomotive, detaching the cage from the rope and allowing the cage to drop 400 m down the shaft. It was believed that had the hook not detached, the rope would have been sufficiently elastic not to have failed, implying that many fatalities may have been avoided. In addition, the risk of accidental opening of hooks due to impact with smaller falling objects in shafts became a concern. This paper deals with a comprehensive statistical investigation that was performed to quantify this risk, establish acceptable levels of such risks and to derive criteria for the resistance of detachment hooks against accidental opening caused by impacts. The investigation involved a survey, dynamic finite element analyses and Monte Carlo simulations.

Jan 1, 2005

By Heyes GW, Orrock GI

The fine grained complex orebody at Elura contains galena, silver and sphalerite in a matrix which is mostly pyrite; sulphides account for about 90% of the orebody. A differential flotation circuit is operated to con- centrate lead and silver in the first section and zinc in the second. The metallurgical performance to date indicates the need for improvement. The lead section has been sampled in detail to comprehensively assess its overall performance and to define problems associated with lead and silver flotation. The substantial amount of pyrite in the feed dominated the perfor- mance of each stage of the separation. It was responsible for extensive dilution of all concentrates and, as a consequence, for very large circulating loads. The rougher concentrate contained 68% pyrite some of which was rejected through the three stages of cleaning to 39% in the final concen- trate. However, this was still sufficient to hold the assay of this product below 45% Pb. Size by size behaviour shows that flotation rates of the valuable minerals, as characterised by their rate coefficients, were low through- out the circuit. There was a significant deterioration in these rates between rougher and cleaner I stages which resulted in large circulating loads between these stages. There was practically no recovery of coarse valu- able minerals to the final concentrate. The dilution of the final concen- trate by non-sulphide gangue was not a major problem. If the recovery of the galena and silver in the +38¦m size range could be raised to that of the - 381Lm size range, galena recovery would rise from 78% to 86% and silver recovery would improve from 46% to 59%. The Pb grade of the final concentrate would remain the same.

Jan 1, 1988

An algorithm is described which leads directly to a unique set of coal washing parameters which, in two or more parallel streams, will give maximum yield of product at a given ash. This occurs when all parallel streams are washed to the same incremental ash. This strategy is shown to offer a considerable increase in yield compared to the simpler strategy of washing each stream to the same sumulative ash.

Jan 1, 1987

By G Chitombo, G Lewis

The knowledge of the expected run of mine (ROM) fragmentation during the feasibility study of a green field site is essential for the design of downstream crushing and milling processes. However,the prediction of blast fragmentation can be a challenging process due to the uncertainty related to the blast design and rock mass parameters. In general,a number of possible blast designs are considered in a feasibility study,with factors such as bench height and blasthole diameter influencing the design. Rock mass parameters are normally derived from a limited number of geotechnical diamond drill cores. Although diamond drill core can provide detailed information concerning intact rock strength (unconfined compressive strength and point load),it provides limited structural information to predict the in situ block size distribution of the rock mass. The in situ block size is an essential input to the prediction of the fragmentation distribution, in particular the coarse end. This paper outlines the methodology of blast fragmentation modeling undertaken for a green field feasibility study at the Riska gold deposit in Indonesia. The favoured milling process for the feasibility study was dump leaching,with no crushing of the ore material extracted from the pit. For this reason,blast fragmentation was a critical issue to be addressed by the study. A range of blast designs were considered with bench heights and blasthole diameters ranging from 4 m to 7 m and 76 mm to 102 mm respectively. Rock mass data was obtained from 19 diamond drill cores across the deposit (total drill length approximately 2200 m). Intact rock strength was estimated from qualitative strength descriptors,while the in situ block size distribution of the rock mass was estimated from the Rock Quality Designation (RQD) of the core. The methodology of fragmentation prediction involved a number of steps including the estimation of rock mass parameters from diamond drillhole data including the estimation of the in situ block size, delineation of blast domains within the deposit,and the assessment of the impact of various blast design parameters and rock mass variation on the predicted fragmentation distribution. The modelling process for the Riska green field site highlighted the need for additional rock mass data to be collected post feasibility stage and once the development of the mine commenced.

Jan 1, 2004

By L E. Setiawan

The mining and milling groups at Batu Hijau are tasked with increasing SAG mill throughput by approximately ten per cent. Avenues used for this process are blasting and crushing optimisation, ore blending and predicting SAG mill throughput in relation to the geological model. This is in line with traditional æMine to MillÆ studies. At Batu Hijau, mill and mine sequencing cannot readily be configured to perform the surveys traditionally required by a æMine to MillÆ study while meeting the daily production target of 140 000 tonnes per day and processing from varying ore types in one pit. A program called BOCCOST (blasting optimisation crushing conveying optimising SAG throughput) has been implemented to capture approximately 200 variables from blasting, mining and milling operations and assign them to a common point in space in the geological model in real time. Interpretive analysis of data from BOCCOST is used to model relationships between the 200 variables to determine which and how these factors affect SAG throughput. Once these relationships are determined, the blasting, ore blending and crushing process can be optimised to generate a product that maximises mill throughput. The relationships modelled from BOCCOST are also used to generate a SAG mill throughput estimator, which can be coded into the production block model. The block model can be reconciled with the geological block model for correlation by geological domain. Once this relationship is established, SAG mill throughput domains can then be established for predicting SAG mill throughput. This SAG mill throughput estimator is then used in conjunction with revenue calculations to determine an ore cut-off, which combines these two elements to maximise metal production.

Jan 1, 2002

In considering the architectural design of the group of buildings which make the surface workings of a mine, the variety of function of the buildings is the most outstanding feature. It falls to the architect to unify the group in a manner which still expresses the function of the individual buildings. In achieving this, many perspective studies were made at an early stage and as the area was developed, a scale model was produced initially for design study purposes and later completed as a model of the final development.

Jan 1, 1957

Stope control information at the C.S.A. mine is obtained by diamond drilling inclined holes from within the stapes. Practical mining outlines are determined from complicated ore contours by a joint mining engineering and geological study. Production holes to achieve this outline are individually marked up and checked. Stope control also extends to selective mucking of ore mixtures for metallurgical reasons and to marginal ore and waste mixtures for economic reasons.

Jan 1, 1969

Western Mining Corporation (WMC), managers of the Windarra Nickel Project, installed a raise-bored haulage shaft and underground crushing system at the Mt Windarra Nickel Mine in 1982. The system is the second of its type installed by WMC, the first being at the Otter Juan Mine in Kambalda in 1976/77. Both haulage systems became necessary as a result of the increasing production limitations, and escalating costs of a deep decline mine.The Windarra haulage system required a single 3 m diameter vertical raise drill hole, 470 m deep and equipped for skip haulage. Two five tonne skips, in balance, hoist crushed ore and mullock to the surface, at a design rate of 200 tonne/hr, effectively reducing truck haulage leads by up to 50 per cent. The system is now fully operational and will enable the mine to achieve its maximum production rate.The raise-bored shaft system has been shown to be cost effective and would be used again in circumstances where raise drilling is applicable.

Jan 1, 1984

Platinum Group metals have been known in New Zealand since the 1880's but not produced on a large scale. They have been neglected because their value was not appreciated rela- tive to gold and their value was often very low, reflecting demand, and because recovery was difficult compared with recovery of gold. With the rising price of most Platinum Group metals and the perfection of gravity and flota- tion recovery methods, recovery of platinum along with gold from alluvial deposits is more attractive. Platinum Group metals are known in several different environments in New Zealand, ranging from alluvial placer deposits to layered basic to ultrabasic intrusions and quartz veins. None of the known occurrences have been ade- quately prospected and there is a real possibi- lity of future discovery of important deposits. Several New Zealand based and at least two offshore resource companies are currently involved in exploration, prospecting or mining for Platinum Group metals in New Zealand.

Jan 1, 1987

Coal mining safety and health legislation is more than typical of safety legislation which has been born of adversity. Indeed, in the main it has been the child of disaster. Requirements which now appear obvious were indeed only occasioned by either social injustice or more commonly by great disasters. Even then changes were often made only after months or years of deliberation by committees. Early mining legislation in New South Wales benefited from the experience in Great Britain. The pioneers of our industry were miners from there and brought with them the basic British legislation, so we started with knowledge gained the hard way. A review of accidents statistics provides a measure of the effectiveness of mine safety legislation, but unfortunately rules cannot control human nature. Miners, managers and inspectors must work as a team to make mines safe and new legislation in New South Wales is planned to this end.

Jan 1, 1975

Traditionally, the term ætotal systemÆ design applied to comminution conjures images of a design that addresses the usual business and operational drivers, including health, safety and environment (HSE), capital cost, operating cost, throughput and product size. In terms of the physical process, the mine to customer type studies also exemplifies an approach that satisfies a normal ætotal systemsÆ approach. As an industry there are drivers that are forcing a re-consideration of what we regard as our total system. This modification of drivers will require changes in approach and implementation to provide the best solutions. Some of the major changes that impact the drivers are shortage of skilled personnel, changing social attitudes to remote mine site work, greenhouse gas emissions (GHGE), quality of available orebodies and geographical location of new mine sites. One approach to the challenges presented by the changing business environment is to expand the system envelope around comminution to ensure that comminution, in its widest sense, helps to meet the new realities. Background to the new drivers and constraints is provided and some of the appropriate responses are examined in relation to the impact on the total system.

Jan 1, 2007

Mount Isa Mines Limited commissioned the Heavy Medium Plant in August 1982. The first processing stage using Mogensen Sizer 2056 dry screens, experienced unresolvable cloth blinding problems in attempting to separate fine, damp Lead/Zinc ore at 1.7 mm. A large proportion of the fines incorrectly reported to oversize and hence to the wet preparation screens. This wet preparation screen overloading produced multiple operating and maintenance problems. The plant configuration precluded retrofitting any alternative proprietary screen. An internal Mount Isa Mines research and development programme was established to design and build a screen to achieve the required per- formance within the site constraints. The resulting ISA SIZER screen has exceeded metallurgical performance expectations. A `clean sheet' approach to the design also permitted elimination of maintenance difficulties experienced with the Mogensen Sizer. Four `production model' ISA SIZERS were installed in February 1988. The ISA SIZER screen, now fully developed is available commercially and will satisfy many screening plant requirements.

Jan 1, 1988