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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

By Thompson M

The humble gabion basket is to mine development what the light bulb and safety pin are to eir-ryday life. Gabions provide a flexible, easy to assemble and cost efficient solution for "101" mine and civil engineering development applications.In some instances gabion baskets are still efficiently in use after many years of faithful, maintenance-free service. In other situations, the baskets have been demolished or made redundant within days, either having served their temporary usefulness or negligently destroyed by the operator of heavy equipment. Either way, the gabion basket must surely rate among the top 10 cost efficient engineering construction materials.

Jan 1, 1994

Permian ultramafic and associated rocks form a belt in East Nelson and contain widespread, but minor, chromite and copper mineralisation and some mining was carried out last century. Since mining ceased, about 1910, little mapping of the belt has been carried out although its origin has been debated vigorously. Recently it has been inferred to form part of a ophiolite sequence, which has been tectonically disrupted to form a melange, and several 'plate -tectonic' models have been proposed. Mapping in the central part of the belt has shown that mTlanges are to the east of the ophiolite sequence. Faulting is widespread so that in anyone section part of the sequence may be absent. Some of the faulting is difficult to recognise and probably took place during or soon after the sequence accumulated. Other faulting is clearly of Cenozoic age and includes major NNE striking faults, with large horizontal displacement, and approximately east-west striking faults with almost entirely vertical movement. The writer agrees that 'plate tectonics' could provide an acceptable explanation for the origin of the ophiolite sequence but the models so far proposed do not take into account the nature and field relationships of the rock units present. Copper Mineralisation is largely controlled by the younger faults, particularly those striking about north-east, and is largely confined to the western part of the ultramafic rocks particularly where slivers of the volcanics of the adjacent Lee River Group are enclosed. Mineralisation, along NNE striking shear zones, has recently been found in the volcanics some distance from the ultramafic rocks. The chromite mineralization is in the form of primary layering in the basal part of the ultramafic belt although in places it has been concentrated into lenses probably during serpentinisation.

Jan 1, 1977

By W Stange

Assessing the risks in a mining project is a key aspect of project evaluation. This is an exacting and demanding task because of the complexity due to the need to deal with uncertainty over long time periods. The way risks are assessed and managed is critical to the long-term success of the projects and the business. Discounted cash flow (DCF) analysis is commonly used for project and risk evaluation. A discount rate is selected that accounts for uncertainties in the cash flows. However, it is difficult to determine an appropriate discount rate due to the large number of factors that need to be considered and the complex relationships between risk sources. Using a proven framework, ævalue at riskÆ (VaR), this paper proposes to rigorously model and quantify ærisk at sourceÆ. The approach was first employed to assess financial market risks and is now widely used in the banking sector for risk assessment. In project evaluation, all risk sources are identified and estimated separately to ensure the overall project risk is quantified accurately. A bottom-up approach has been proposed. Estimates are first carried out for risk factors at the bottom of the project structure. The assessment is then applied at increasingly higher levels until the overall project risk is estimated. Value at risk is a versatile methodology and can be used in a local or integrated manner. The final value is objective and the calculation steps can be easily justified and followed. Four case studies have been included to illustrate how the proposed approach can assist management in decision-making.

Jan 1, 2009

By R W. Wilson, A R. Branch

For hot mines, the use of mechanical refrigeration machines has become an essential feature of ventilation and cooling systems. Over the last 30 years, considerable effort has been made in developing refrigeration strategies and equipment to provide cooling that is cost-effective, reliable and energy efficient. In addition to these general design requirements it is necessary to take into account project specific factors such as long term mine cooling needs, operating strategy and installation time. This paper examines a number of actual mine cooling installations which illustrate how these features have influenced the design and installations. For many mines, it is known how the cooling requirements will increase as the mine develops, based on the life of mine plan. The design of the cooling installation must cater for both short- and long-term needs. The initial installation (to meet short-term requirements) must form an integral part of the ultimate cooling installation and as additional equipment and refrigeration machines are installed, it is necessary to ensure that the system remains cost-effective, reliable and energy efficient at all stages. For many mines, refrigeration machines are installed on surface and either operated near-continuously, or for the summer months only. Different maintenance strategies and policies with regard to standby equipment are highlighted for three mine installations. In spite of developments in cooling strategy, the cost of providing refrigeration remains expensive and many mines delay the introduction of refrigeration, as long as possible. When it is finally accepted that refrigeration is absolutely necessary, it is frequently necessary to æfast trackÆ the installation. To minimise site erection time and meet installation deadlines, factory pre-assembly and standardised modular equipment are widely used.

Jan 1, 2005

My closest personal connection with the problems of rehabilitation comes from two years spent on a project at Weipa. At the beginning I am fairly certain that the engineers did not understand the mental processes of the ecologist they were dealing with. I am absolutely certain that the ecologist did not understand the minds of the engineers. This form of the two cultures problem is the root of some of the matters that lie behind this lecture. It is possible to establish a stabilising cover of vegetation on virtually any land surface, short of a cliff. While the biologist knows this he also knows that the physical and chemical nature of a given acre of land combines with climate to define quite sharply just what plants can be grown there. In the case of land after mining the range of plants may be quite limited. The engineer, and his masters in the accounting section, seem (in my experience) to expect that rehabilitation can be carried on with equal speed and facility in any location. This belief commonly leads to a situation where rehabilitation is not considered until mining is well under way. True, the mine must be proved as an assured profit- earner before money can be justified for the non-productive (to the miner) expense of rehabilitation. But when the miner finally starts rehabilitation he does not always get a convincing and speedy result. Furthermore, the failure may not be easy to overcome. This is even more likely to be the sequence of events if the miner wants, or is obliged, to specify the exact type of vegetation to be produced, e. g. commercial forest, pasture,' or the original native vegetation.

Jan 1, 1971

By J D. van Wees

Preliminary project evaluation is relevant to the evaluation process of the decision makers (or executives) prior to making the critical business choices. Important to the assessment is an integrated sensitivity, scenario and risk analysis at every stage of a mining project. This should include the evaluation of different options taking into account uncertainties. Construction of interactive financial computer models where each element is functionally integrated into a fully integrated life-cycle allows rapid and automatic study of the sensitivity of each element through changes in the value of an input. The scenario analysis is the effect of changing more than one input at a time, organised through scenarios in a decision tree. Risk analysis is an integrated analysis of the effect of all uncertainties of the parameters, using Monte Carlo simulation. Key to the asset decisions process is that the decision maker is provided with the essential integrated decision information. We have clustered this information into several scenario groups. Within each group, two levels of uncertain information are distinguished: discrete scenarios critical for the decisions; and remaining uncertainty within the scenario (risk analysis) which can be modelled by continuous probability density functions (pdfÆs). For an asset, the relevant scenarios in each group and the associated continuous pdfÆs need to be defined. When the scenario groups and associated pdfÆs have been defined for the asset, the decision process can be started by structuring the different scenario groups in a Scenario Tree. A Scenario Tree contains decision and end nodes. Each end node corresponds to a completely defined mine project or asset model. The calculations for each end node result in a set of stochastic realisations for the reserves, production forecast, cash flow, and profitability indicators (NPV, IRR, DPR). Thus, a deterministic and fully probabilistic multi-commodity analysis tool is provided to investigate the sensitivity and robustness of decisions.

Jan 1, 2003

By M P. Sandy, J B. Simpson

The empirical stope design techniques developed by Mathews et al (1981) and refined by Potvin (1988) and others, have gained rapid acceptance in Australia in the 1990s as a simple, æfirst-passÆ means of designing primary stopes. However, it is recognised that these æstability graphÆ methods can only provide broad design guidelines. Recently published stability graphs (Hutchinson and Diederichs, 1996) encompass a wide range of geological environments, which are not discussed, and the influence of time and blasting practices cannot be easily reviewed. The æscatterÆ in the graphs must reflect, to some degree, different operating practices as well as subtle but important differences in the pre-mining conditions that are not adequately accounted for in the æmodified stability numberÆ. Development of a ælocalÆ, site-specific stability graph may overcome some of these problems. At NormandyÆs Golden Grove Operations, Western Australia, detailed reviews of stope stability performance, including Cavity Monitoring System (CMS) surveys, are undertaken for every stope, with the objective of developing a series of local stability graphs, customised to site conditions and calibrated in terms of æAverage Falloff ThicknessÆ (AFT). Cablebolting of stope hangingwalls is used to develop reinforced æbeamsÆ on each sublevel, effectively dividing the full span into smaller, more stable æsubspansÆ. By analysing stope stability performance in terms of fall-off, cost-benefit studies can be undertaken of the effect of cablebolting and other options for modifying stope spans. The approach used to develop the local stability graphs, and their integration into routine stope design procedures at Golden Grove, is described. The experience gained during stope extraction in the Scuddles orebodies has been successfully applied to the Gossan Hill æC-PanelÆ orebody. This methodology should be applicable to other underground mining operations.

Jan 1, 2000

By T Suwa, S Ishida, J Xue, S Fuchida, H Shimada, Y Aoki, C Tokoro

Pyrite depression with prolonged pre-conditioning treatment in high salinity water has attracted much of researchers’ interest in recent years considering the increased utilisation of multiple stage flotation, closed water circuits, clay minerals, and sea water in flotation plants. In this study, the role of calcium (Ca)/magnesium (Mg) precipitates in stabilising the dissociative ferric hydroxy-xanthate (FHX), and its effect on modifying the balance of hydrophobic/hydrophilic species on the pyrite surface were primarily investigated. The total amount of extractable Fe-bearing species on pyrite concentrates and tailings after floatability test was quantitively determined using ethylenediaminetetraacetic acid (EDTA) extraction tests combined with inductively coupled plasma optical emission spectroscopy (ICP-OES). A more evident accumulation of extracted-Fe in pyrite concentrates rather than it in pyrite tailings was observed at both pH 9 and 10.5 when pyrite was pre-conditioned in Ca/Mg-bearing solutions for 360 min. Fourier transform infrared (FTIR) spectroscopy showed the main xanthate derivatives changed from di-xanthogen (1155 cm-1 ) to FHX (1167 cm-1 ) on the oxidised pyrite samples that were preconditioned in Ca/Mg-bearing solutions. Scanning electron microscopy-energy dispersive spectrometry (SEM-EDS) suggested the preferential accumulation of FHX upon secondary Ca/Mg-bearing precipitates as reflected by the higher Fe/S ratio together with the decrease of O element herein after xanthate conditioning. Although the hydrophilic secondary Ca/Mg/Fe-bearing minerals on the pyrite surface generally deteriorate pyrite’s floatability, the upper section of these secondary hydrophilic aggregates is partially covered by FHX alongside with further multilayer-adsorbed di-xanthogen. The precipitated Ca/Mg-bearing precipitates promotes the immobilisation of dissociated FHX derived from the reaction among iron oxyhydroxides and xanthate anions upon the localised secondary Ca/Mg/Fe-bearing minerals via acid–base complexation. Therefore, this phenomenon alleviates the hydrophilic modification of secondary Ca/Mg/Fe-bearing minerals to pyrite considering the immobilised FHX upon localised precipitates shows a slight hydrophobicity, and meanwhile, the upper FHX benefits the multilayer adsorption of di-xanthogen via the interaction among their inner hydrophobic alkyl chains.

Aug 24, 2022

Pasminco is actively addressing the optimal utilisation of the known orebody at the Broken Hill mine, in addition to actively exploring for economic mineralisation in the region. As part of the policy of better utilising the known mineralisation a major upgrade of geological systems at the mine was initiated. The software chosen for this change was Maptek's Vulcan, mainly because of the ease of use of this package. Rather than having numerous data sources in varying software packages the mine aimed to limit interfaces and packages. The decision to upgrade the mine's software was not taken lightly, especially in view of the inature stage of Broken Hill mining operations. However, the current mine life of Pasminco Broken Hill mine is I I years, which is longer than many new mines setting up in Australia today. Also, a more sophisticated system brings with it a better understanding of the economic mineralisation, possibly extending the mine life further. The process of moving the Broken Hill databases into Vulcan was a huge task which included, database transfer, digitising, wire framing, modelling and eventually ore resource calculation. To ensure the validity of the new ore resource accounting system, the existing system was run as a duplicate for the introductory year, providing a comparison and if necessary a safety net. The benefits of such a change, however, became apparent almost immediately. Area geologists realised the opportunity to revise geological interpretations in their areas of responsibility, enabling them to obtain a better understanding of orebody morphology. Three- diniensional analysis provides an excellent method by which review of ore targets both in the underground and mine lease exploration areas can he made. As each new find of mineralisation enters the system and hopefully extends the mine life, the opportunities for further mineral exploration in Broken Hill area by Pasminco are enhanced.

Jan 1, 1997

Agitation leaching data has generally not been analysed to extract the maximum amount of information that is available. Analysis of heap leaching systems has been developing over the last five years to include the underlying rate determining process. This has been found to be diffusion control for all copper leaching; irrespective of whether they are oxide or sulfide minerals. Extension of the diffusion model has enabled better scale-up protocols to be developed. Acid consumption in copper heap leaching operations has been found to be a first-order chemical controlled system. The analysis of the data is relatively straightforward. However the interpretation and scale-up is not as direct as for agitation leaching. The differences in the residence time distributions (RTDs) are required to properly account for the scale-up between test columns and field. The leaching of oxide copper in agitated systems has also been found to be diffusion controlled. However, there are significant differences in the control mechanisms. Initially the leaching control is the supply of acid to the mineral surface. This is limited by the diffusion through the surface boundary layer and is thus subject to the level of agitation. Higher agitation provides thinner boundary layers and faster diffusion rates. Similarly, higher grade material creates a thicker concentration boundary layer and relatively slower leaching. Acid consumption in agitated leach systems is also a first-order system. However, the interpretation of the RTD is easier from batch testing to continuous multiple CSTR series plant operations. Altering the method of acid delivery can significantly reduce acid consumption rates. Predilution of the acid minimises the local acid concentration near the addition point and can reduce the consumption by up to 50 per cent. Modelling of the combined leach rate(s) and acid consumption rates in agitated systems allows comparison of various leaching strategies and selection of the most appropriate. Similarly, the model allows the identification of economic limits to leach recovery; where the cost of acid consumption is greater than the value of the copper recovered. From this analysis selection of residence times for agitation leaching can be determined. The modelling tools are useful in identifying optimum conditions for the leaching process.

Jan 1, 2005

By M Byrne, P Newman, M Lindqvist

SomincorÆs Neves Corvo Mine, one of the largest copper producers in Europe, is currently constructing a paste backfill plant in order to replace cemented rockfill in its bench and fill stopes and possibly replace hydraulic sandfill in the drift and fill stopes. A 250 m flow loop was constructed at the mine and the flow properties of various paste blends were investigated. Cycloned tailings and total tailings were both pumped successfully as pastes and were also combined with waste rock for additional testing. The pressure losses in 102 mm and 154 mm diameter pipeline were recorded at flowrates up to 60 m3/hr and are presented at slumps from 120 mm to 250 mm. The data showed the sensitivity of paste mixtures to slight changes in consistency, confirming the need for accurate control during paste production and distribution. The flow results clearly showed the different paste behaviour between cycloned and total tailings pastes and the change in properties at low and high slumps. In addition to pressure loss data, strength testing was done on 38 mm and 82 mm diameter samples. Further strength testing is ongoing at the mine using fly ash as a supplement to cement, reducing the overall binder cost. Preliminary engineering for the underground distribution system is briefly discussed, with reference to the siting of the plant. The Neves Corvo Mine is particularly shallow and alternative methods to pumping were sought in order to ensure sufficient hydraulic head for underground distribution. The comparison of options had to take into account the capital and operating costs, the operation of the plant and the impact on the future expansion of paste fill utilisation at Neves Corvo.

Jan 1, 1998

By Y Potvin

Risk assessment in the mining industry is currently a topic of great interest with risk methodologies being increasingly applied beyond the traditional areas of finance and feasibility to analyses of risks due to various physical hazards including chemical, electrical, mechanical and geotechnical sources. The application of risk analysis and evaluation techniques to the management of personnel safety is being promoted by both industry and regulators but, particularly for geotechnical hazards, often there are inadequate databases for a meaningful quantitative calculation of risk. In Western Australia over the decade 1987 - 1997, the annual fatality rate for underground mining has averaged at approximately 1.3 in 1000 employees with rockfalls/rockbursts accounting for over half of these deaths (Prevention of Mining Fatalities Taskforce, 1997). Recent improvements have been made since this time, perhaps halving the annual fatality rate, mainly due to industry improving ground control measures and implementing various recommendations from the Prevention of Mining Fatalities Taskforce (1997). However there still remain a large number of unknowns when trying to address the complex area of geotechnical risk in mining. Risk calculations require information about the source of the hazard, the people or assets exposed to the hazard (the æelements at riskÆ) and the effects on these vulnerable elements if the hazard actually occurs (the potential consequences). To date, little information has been available on the exposure of mining personnel and equipment to underground hazards other than occupational safety studies on chemicals, heat, dust, etc. This paper concentrates on underground geotechnical risks and presents a data collection and rating system for the exposure of both the underground workforce and equipment to rockfalls and rockbursts. The methodology considers exposure times, personnel protection levels and hazard proximities to give exposure ratings for the major underground mining activities and hence rate different locations within a mine. This exposure methodology is then applied to the workforce and production/ development cycles in an actual WA underground mine, creating a snapshot map of differing levels of personnel exposure across the mine.

Jan 1, 2003

By D Thornton

Geologists use large data sets and spend many hours and much computing time modelling the distribution of minerals throughout a rock mass. Significant investments in drilling, assaying and quality assurance and quality control (QA/QC) are made before sophisticated geostatistical techniques are applied to the collected data. A large amount of research has been done in this field such that the ore control model is a reasonably accurate reflection of the mineral that is actually in the ground. However, after all this effort, the mining engineers then work just as diligently to blast the in situ rock mass to a particle size that can be efficiently handled and further reduced by downstream comminution processes. In open cut mines, this invariably requires large amounts of explosive energy that displaces the ore boundaries from where the geologists originally defined them. Failure to accurately account for blast movement will result in misclassification, whether that is ore to waste; low grade to high grade; sulfide to oxide; or other contaminates û collectively referred to as ore loss and dilution. The financial consequence of getting this wrong is substantial and not well quantified. This diminishes all of the good geological QA/QC done in defining the ore. This paper discusses various technologies used to physically measure blast-induced orebody movement and the difficulties in modelling this with a purely theoretical approach. Rock mass structure, blast energy, design and timing contours all have significant and interrelated effects on blast movement vectors. Measured blast movement has a large variance due to the uncertainly of many of the controlling parameters û arguably dominated by the heterogeneous nature of the rock mass. Although there have been significant advances in blast movement modelling capability in recent years, this unpredictable blast movement limits their use in the context of production ore control. Certainly, without physical field measurements, accurate predictions of movement are near impossible.Limitations of traditional surface markers in determining three-dimensional (3D) blast movement within the orebody are also reviewed, and advancements in accurately measuring these vectors within blasted ore discussed

Sep 26, 2011

By D S. Scovira

The effect of blast vibration on highwall stability is a common concern for many open pit mines. Vibration in such a case is a typical near-field blast vibration situation. Vibration modelling in the near-field of a blast requires consideration of both the amplitude and frequency attenuation of waveforms with distance. Most existing blast vibration models are designed for far-field vibration and do not simulate the frequency attenuation and its effect on waveform shape. With the advent of electronic detonators it is desirable to quantify the effects of blast delay on the amplitude as well as the frequency of blast vibrations. A suitable modelling tool is proposed here. At present, a blast vibration model with double attenuations(amplitude and frequency) has been developed at Orica. The present model simulates the amplitude and frequency attenuations separately using different transfer functions. The amplitude attenuation is modelled according to the non-linear relation between the peak particle velocity (PPV) and the charge weight scaled distance established from the signature hole vibration. The frequency attenuation is modelled with the signal broadening based on KjartanssonÆs transfer function û a linear model for attenuation of waves. The model is capable of taking multiple sets of signature waveforms as input to improve the modelling accuracy of the wave propagation in the geology at a site. The screening effect of the earlier firing holes in the path of a later firing hole is taken into account on not only the amplitude, but also frequency attenuation. The present blast vibration model is suitable for modelling both near-field and far-field blast vibrations. It yields reasonable agreement with measured field data for the near-field vibration examples discussed.

Jan 1, 2007

By Denby B

The ever increasing demand on land use has led to large areas of land being considered for development, which because of past mining activities are covered by substantial depths of fill materials. This paper examines the problem faced with the development of land infilled with open cast mining backfill, in predicting the behaviour of the fill upon development. To be able to design foundations suitable to withstand the movements that occur within open cast backfills or design a backfilling operation that produces land suitable for a proposed after development, a means of predicting backfill settlement is required. Factors critical to this prediction are the timing of backfilling operations, the compactive state of the backfill and the inundation of the backfill. The analysis of a considerable quantity of data collected from the monitoring of backfilling operations and backfill movement at a range of open cast coal sites located within the UK has given the authors a better understanding of how open cast backfills behave in terms of long-term creep settlement, collapse settlement and how these are effected by backfill scheduling. This knowledge has been encoded into a program running under Microsoft Windows that enables a user to make settlement predictions for a point, a given section or area or for the whole site. Differential settlements, critical to any development design, can also be determined between any two points taking into consideration the heterogeneous nature of open cast backfills. The program was tested with data from an open cast coal site in the UK. Compaction monitoring during the placement of the backfill material enabled a model of the backfill to be constructed taking into account backfill properties and scheduling. This model, taken as input into the program, enabled future backfill settlement predictions to be made. A comparison is given between these results and actual settlements obtained from a short-term settlement monitoring scheme.

Jan 1, 1995

By G Milligan

This paper describes how water is being planned, managed and used in Queensland as both a precious natural resource, vital for healthy ecosystems, and as a valuable tradeable commodity, necessary to secure economic prosperity. Queensland has been progressively implementing a planned package of water industry improvements agreed by the Council of Australia Governments in 1994. The Water Act 2000 provides the legislative basis for implementing the COAG water reform framework in Queensland and sets out statutory water planning processes. A Water Resource Plan (WRP) is a strategic document that specifies water sharing and environmental flow outcomes across a whole catchment over a period of up to ten years. A Resource Operations Plan (ROP) sets out the details of implementing a WRP, including environmental flow rules, establishment of tradeable water allocations that are separate from land, rules for water trading, rules for operating water infrastructure, arrangements for the release of any unallocated water identified in the WRP, and requirements for monitoring and reporting. Water Resource Planning and Resource Operations Planning are progressing in priority catchments and aquifers across the State. Ways to obtain a water entitlement now depend on whether a WRP and/or a ROP exist in a particular catchment. Water trading provides for the buying and selling of water entitlements on a permanent basis where permitted under a ROP. Details of tradeable water allocations are securely registered on the Water Allocations Register, which is similar to a land titles register. Leasing and temporary transfers of water entitlements are also possible. Water brokers are emerging to facilitate water trading which aims to improve water use efficiency and move water to higher value uses. Queensland is leading the way in the implementation of the COAG water reform framework and the pursuit of sustainable water resource management. The StateÆs progress in this regard is increasing certainty for water users, while ensuring that considerations of equity amongst existing and future users and amongst various uses, as well as considerations of ecological sustainability and the wider community, are all taken into account. The mining sector stands to benefit substantially from these new approaches to planning, managing and using this precious resource.

Jan 1, 2003

By J Vann, S Dunham, M Job

Two-dimensional (2D) estimation approaches have significant advantages over three-dimensional (3D) approaches when dealing with a narrow vein- or layer-like geometry. The main motivation for 2D estimation is that bias due to non-additivity of grade is avoided. The various aspects of 2D estimation are illustrated by use of a case study from the 1A Shoot Nickel deposit at Leinster, Western Australia. The width of the 1A mineralisation is variable, and so are the angles of drilling intersection. Consequently, the variable of interest (in this case nickel grade) is not suitable for direct kriging because it is defined on non-uniform support. However, grade can also be defined as the ratio of two variables that are suitable for direct kriging: thickness and accumulation. In the 2D approach, tonnage equates to thickness and metal equates to the product of grade and thickness, also called æaccumulationÆ. In the simplest approach we can independently krige accumulation and thickness and derive the grade by the ratio of the two resultant estimates. The thickness and accumulation variables must be in a consistent projection orientation, for example a horizontal projection in the case of steeply inclined mineralisation like 1A. Because the density is not constant across the 1A (and most other) deposits, the issue of density weighting is also discussed. The impact of the correlation between grade and length, or grade and density are important issues in implementation of 2D estimation. The circumstances under which these correlations need to be taken into account (by co-kriging) are reviewed. Finally some comments are made on the final use of the resultant 2D estimate for mine planning purposes.

Jan 1, 2003

This paper summarises the performance of the gold exploration industry in the Western World since 1970, with particular emphasis on the period 1985 - 2003. It identifies the number and size of major (ie >1 million ounce) primary gold discoveries made over this period by location and type of company. Estimates are also made of the likely contribution from smaller discoveries down to 0.1 million ounces. It also assesses the overall trend in the percentage of grassroot discoveries that are converted into mines, and the time taken to get them into production. By comparing the discovery data against total exploration expenditures it is possible to determine the average unit cost of finding a major gold deposit. It is significant to note that the average unit cost per discovery and per ounce of resource has tripled over the last 30 years. An attempt is made to determine the overall economic value generated by the industry. This is based on a detailed analysis of the rates of return and the Net Present Values (NPVs) of 109 major gold projects. After including an adjustment for discoveries not modelled, the NPVs are compared against the total amount of money spent on exploration to determine the value-to-cost ratio for the industry. Based on historical spot gold prices and tax rules, it is found that the return on investment for gold exploration in the Western World over the period 1985 - 2003 was around five per cent real after-tax. The returns achieved in the US and Australia was significantly higher than that for Canada. The returns also appear to have declined in recent years. With regard to the relative performance of junior versus major companies, the paper estimates that the junior sector accounted for 30 - 40 per cent of the total exploration expenditure, and were was associated with 34 per cent of the total number of major deposits, 27 per cent of the contained ounces and 25 per cent of the economic value found.

Jan 1, 2004

By J N. Moncrieff

The resource estimation process for the underground Mount Isa Copper Mine has undergone significant change. Block models are now created using a commercial software package as opposed to the in-house software that was previously used. With mineralisation occurring over a strike length of more than 4 km, Mineral Resources for all categories totalled 190 Mt at 2.7 per cent Cu at 31 December 2002. Estimation is by ordinary kriging with two separate models generated to cover the X41 and Enterprise orebodies respectively. The underlying Upper Basement Contact and the silica-dolomite alteration envelope are both used as hard boundaries to constrain the modelling. From a validated block model, grade shells are generated at the appropriate cut-off grade for a given area. The grade shells are depleted for mined voids and subdivided into 30 metre strike length blocks. Application of a SMU criteria and crown and pillar constraints results in failed blocks being culled from the resource inventory. Mineral Resource blocks are further subdivided into resource categories (Measured, Indicated, Inferred) by the Competent Person on the basis of geological confidence, taking account of geological and grade continuity, drillhole spacing and structural complexity (JORC, 1999). A geostatistical æconfidenceÆ category is incorporated into the block model as an additional tool for the categorisation process. The shapes are assessed in concert with a mining engineer to determine whether a reasonable expectation of mining potential exists and if not they are removed from the resource. The final categorised blocks constitute the Mineral Resource inventory and form the basis for conversion to Ore Reserves. The resultant Mineral Resources and Ore Reserves are reported on an æinclusive ofÆ basis.

Jan 1, 2003